The metallurgy of the common metals, gold, silver, iron (and steel), copper, lead and zinc
This outline of the metallurgy of the common metals, namely, gold, silver, iron, copper, lead, and zinc, is devoted to the description of processes for
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The metallurgy of the common metals, gold, silver, iron (and steel), copper, lead and zinc is a 1921 historical mining reference by Austin, Leonard S. (Leonard Strong), preserved in the Mountain Man Mining research library. This outline of the metallurgy of the common metals, namely, gold, silver, iron, copper, lead, and zinc, is devoted to the description of processes for...
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The Metallurgy Of The Common Metals
Leonard S. Austin
Formftiy Pntfettor of MdaUurgj/ and Or* Dnttutg, MiMgan Cclkgt of Mina
Fifth Edition
New York
JOHN WILEY & SONS. Inc.
London: CHAPMAN k HALL, Lootbo
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Wv-T."-
HARVARO UNIVERSITY EMINEeRlNQ SCHOOL
Copiriqht
Edition, AltNiHa & Scientific I
Third Fourth
Fifth
Lkonasd S Adstin
Bs, 1907 1909 1011 1913 1921
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Preface To The Fifth Edition
Since 1913, the d&te of the last edition, Buch ntdical changes and improvements have been made in the metallui^ of the common metals, that this edition of 1921 has been largely rewritten to bring it in accord with present practice, as will be seen by examination of the following pages. Great pains have been taken to dearly set forth underlying principles and at the same time to give the details of methods and of metallurgical equipment, and their cost. It is reaUzed that, due to the rapid advance in prices, the costs of operation have lately been subject to serious modification. A chapter has been devoted to questions of the economic situation (A the business of metallurgy. Little attempt has been made to describe methods not now in use.
L. S. Austin.
Lob AnoeIjE8, May 1, 1921.
Preface To The First Edition
This outline of the metallurgy of the common metals, namely, gold, silver, iron, copper, lead, and zinc, is devoted to the description of processes for winning these metals from their ores and then refining them. The metallurgy of iron is treated only to the point where pig-iron is obtained.
Following the description of ores, as well as of the fuels used in smelting tbem, and the materials of which the furnaces are constructed, we come to the sampling, fcv the determination of the exact value of the ore before treatment.
A chapter has been devoted to the subject of thermochemistry as applied to igneous methods of extraction. The winning or reduction of the various metals is then taken up in order, and is followed by a description ol the methods of refining them. Attention is then given to commercial considerations, since the processes must be conducted in a profitable way.
The author is indebted to Mr. F. L. Bosqui, who has not only read the manuscript, but has modified the portion devoted to the cyaniding of gold and silver ores, as his special knowledge' has justified. For the subject matter relating to the smelting of silver-lead and copper ores, the
author baa drawn on hia own experience, ^lined during a quarter of a century of practical work.
L. S. Austin. HoDQHTON, May 1, 1907.
Preface To The Second Edition
The experience gained in using the first edition has suggested many changes, and the book has accordingly been re-written, adding new matter, describing other processes, and keeping step with modem practice.
In Part I the subject of thermo-chemistry has been expanded, and a table of heats of formation given. The description of the cyanide process has been amplified and brought up to date, for miUing methods are being rapidly improved, and cyanidation is having increased application, especially in the treatment of silver-bearing ores. The metallurgy of zinc has been treated more fully, and particular attention given to the principles underlying the smelting of zinc ores.
In the part devoted to refining there has been added the making of wrought-iron and steel, the refining of zinc, and the electrolytic refining of lead.
Plant and equipment is placed in a separate chapter, while the division describing the economics of metallurgy has been thrown into a more systematic form. The author is indebted to Mr. E. A. Hersam, who read the manuscript of the second edition and made numerous su^estions and corrections.
The author is indebted to the following companies for the use of certain of the illustrations in this book: AUis-Chalmers Co., MUwaukee, Wis.; Power 4 Minii^ Machinery Co., Cudahy, Wis.; Chieholm, Matthew & Co., Colorado Springs, Colo.; F. M. Davis Iron Works Co., Denver, Colo.; Steama-Roger Mfg. Co., Denver, Colo.; Pacific Tank Co., San Francisco, Cal.; Redwood Manufacturers Co., San Francisco, Cal.; Galigher Machinery Co., Salt Lake City, Utah; Traylor Engineering Co., Allentown, Pa.; Blaisdell Co., Los Angeles, Cal.; Denver Engineering Works Co., Denver, Colo.; Treat Engineering* Machinery Co., Salt Lake City, Utah; Risdon Iron Works, San Francisco, Cal.; Colorado Iron Works Co., Denver, C<do.; Cyanide Plant Supply Co., Ltd., London; The Jeffrey Mfg. Co., Columbus, Ohio.
L, S. AuBTm. HouoirroN, August 1, 1900.
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Preface To The Third Edition
The preaent edition has been more systematically arranged, and error have been eliminated. Important and recent changes in smelting practice and in the cyanidation of gold and silver ores have justified the insertion of additional matter, much of which has come under the direct observation and inquiries of the author.
L. S. Austin. Salt Lake Gitt, Utah, M&rch 1, 1911.
Preface To The Fourth Edition
In this edition, that part of Chapters X and XIV which discusaes the cyaniding of gold and silver ores respectively, has been written by M. W. von Bemewitz, formerly of the Associated Northern and Associated Mines, Kalgoorlie, Western Australia, and now on the staff of the Mining and ScienUJu: Press. The chapter on the metallurgy of zinc has been re-written by Mr. R. G. Hall, long manager for the United Zinc & Chemical Co., and later in general consulting practice. These gentlemen are s)ecially qualified for the subjects they have undertaken and have incorporated the recent practice in the art.
L. S. Austin, ' Salt Lake Cnr, Utah, August 15, 1913.
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Contents
Part I— General Metallurgy
CHAPTER I Orbs aks Mbtaia:
Definition and CbssifieatioD of One 3
Methods of Treatment 4
Claoaificatioa of Met&Uurgicd Operations 5
Principiea Relating to the Refining of Metob 5
IkiouldinA and Caatiiu Metals . ^ .,.. ^ ..... ^ ,, ^ ...... ^ .,,,.,..,... ^ ... . 7
CHAPTER II Fuels:
Pudi 11
Hw Naturri Solid Fuels 12
Producer Gas 33
Pulverued Coal 29
Chapter Iii
Rcfbactoriks:
Refractory Materials and Their Properties 31
Acid Refractmies 32
Neub«l and Basic Refractories 37
CHAPTER IV Th» Prepahat.on or Ores:
Principles of Sampling 39
Receiving, sampling, crushing, bedding and storing ores 40
Sampling Metals 48
CHAPTER V CauBRiNa, Gmndino, Screenino and Ct^seimNa:
Principles of Crushing 60
Coarae or Primary Crushing 51
Intermediate or Fine Crushing or Coarse Grinding 66
Fine Grioding 64
Screening 66
Claenfyine 68
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viii CONTENTS
CSAPTER VI Mbtallttroical Furnaces:
Shaft Furnace ; . . 72
Reverbenitory Fumace 73
Chapter Vii
Combdstion :
Temperaturefl of Combustion 79
CHAPTEE VIII Mbtalldbuical TBERUO-caEUisTBr:
Methods of Determining Thermic Valuee S4
Heate of Formalion of Chemical EllementB 86
CHAPTER IX RoABTiNo:
Kinds of Roasting 8S
Chemistry of Roasting 89
Roasting Ores in Lump Form; Heap-roasting 92
Roasting of Ores in Pulverised Condition 94
The Long-hearth Reverberatory Roaster 95
Me«hanically-operated Roasting Furnaces 97
Roasting trf Matte 108
LoB8e« in Roasting 108
Capacity of Furnaces and Cost of Roasting 109
Blast or Pot-roasting UO
Sinter Roasting 110
Triple Roasting 113
Chapter X
CONCBNTBATION OF OrES AS A SuBSIDIAfir OPERATION rN MbTAIXUROT:
Concentration 114
Flotation 116
PART II— GOLD CHAPTER XI Gold Ores ano Classification for Millinq:
Occurrence 121
Valuation of Gold and Silver 122
ClaasiScation for Milling 122
CHAPTER XII Aualqamation:
Stamp Milling with Plate Amalgamation 124
General Arrangement of a Gold Stamp-mill 130
Concentration in Stamp-milling 131
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CONTENTS be
Chapter Xiu
HrDROMSTALLimoT or Gold Obxs:
Milling Ores in Aqueous SolutioD 133
Chapter Xiv
Chlobimaiion or Gold Okbs:
Dm Suited to CUoriiMtion 135
TTw Goldfield Chlorine MiU Co., Goldfield, Nev 136
Bftirei CUoriiuttiou 137
CHAPTER XV CiAMiDDia or Gold Obks:
Outlioe of tbe ProMos erf CyaDidiog 143
Ores for CyssidstioD 145
Cbemistrr of the Cyanide Process for OoM Ores 146
lite StAndaid Systems of Cysniding 150
Double TicKtnmt 158
BUme-Agitstion 157
Pneumstic Agitstora 158
Machsnicsl AgitattXB '... 159
Ctoubiued Mectuuiical and Puaiunatic AgitatoTS 160
Agitation Treatment 161
Continuous Countar-currettt DecantAtion 165
FiltzatioD or Sepuation of Metal-bearing Biriutim from Slime 166
Vacuum Filtration 168
Pnsmire Filtration 171
General Remarics on KHers 17S
Tlie Crowe Vacuum Process 177
"nie Precipitation of Gold from Cyanide Solutions 177
Hw Merrill Precipitation Process 178
The Zinc or Extiwtor Box 181
TTw Clean-up 182
Dryiag and Refining the Gold Precipitate 184
Capital Costs <rf Slime Plants 186
CHAPTER XVI TrPiCAi- QoLD-MiLi. PBAcnci:
Cyaniding Free-milling Porous Ores 18B
The Wasp No. 2 Mill, South Dakota 189
Ores of Clayey Nature by Cyaniding 189
The Victorious Mill, Western Austeslia 190
The Kolar Field 191
The City-Jeep Mill 191
Tlie MiU of the Consolidated Umglaaghte Co., Rand, South Africa 193
The Homestake Mill, Lotd, South Dakota 19S
Hie liberty BeU MiU, TeUuride, Colo 199
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Treatment ot Telluride Ores 199
The Golden Cycle Mill 201
The Victor Plant of the Portland Gold Mining Co 202
The Kalgoorlie District, Western AuBtrali& 203
The Oroya^Brownhill Mill, Kalgoorik Diatriot 204
The Hollinger Mill, Porcupine District, Ontario 205
The Tom Seed Mill, Oatman, Aril 207
The United Eastern Mill, Oatman, Aria 209
CHAPTER XVII Tbeathekt or Gold-hill CottcsiniuTEs:
Oasaification 213
The Alaska-Treadwell Concentrate Treatment Pltmt. 215
Concentrate Treatment at the Goldfield Consolidated, GoldBeld, Nev 218
CHAPTER XVIU Vasioub Triatmbntb and Calculations:
Flotation and Cyaniding 223
Drying and Cyaniding 223
Treatment of TailingH from Acid or Ammonia Leaching 223
Calculation of Tonnages in Mills 223
CHAPTER XIX SuKi/rtNO Gold Obxs:
Blast-funiace-smelting m. Cyaniding of Gold Ores 236
The Price <rf Gdd Ores. Also the Cost of I^inducing and Sellii^. The
Price of Gold 227
PART III— SILVER CHAPTER XX SiLVXB, ITS Obxs and Their Treatment:
Characteristics of SUver Ores 231
Extraction of Silver from Ores 232
Treatnient of SUver Oreo 232
CHAPTER XXI Amalgamation op Silver Ores:
Wet Silver Milling with Tank-settling 235
The Boss Ptoceae of Silver Milling 243
The High-grade Nipissing Mill, Cobalt, Ontario 243
Plate and Pan- Amalgamation and Concentratioo of Silver Ores 244
The Chloridiiing Roasting of Silver Ores 246
The Patio Process 249
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CONTENTS n
Chapter Xxii
The PrincipleB of the HydrameUUurKy irf Silver 2S0
The AuguBtin Procen 250
The Ziervogel Ptoccm 261
The Hypomilphite linviatioii Ptocbm for Silver Ores 254
The Ruflaell Piocees 251
CHAPTER XXIU CTANtDATtOK or SiLvza Okes:
nincipleB of Cyanid&tion 256
Ttie Preoipitation of Silver from Cyanide Solution 2S7
Tt>e SantA Gertrudia Precipitating and Refining Pltutt 250
I^edpitAtion by Aluminum Dust 201
Drying and Refining Silver Precipitate 261
Chemistry of the Ptocem tor Silver Oree 264
TVpical Silver Milk 26S
Behnoot Milling Co.'h Mill, Tonopah, Nev 266
Cytmidation of Mixed Silver Ores at the San Francisco Mill, Paehuca, Mex. 268
The Waihi Gnnd Junction Mill, Waihi, N. Z 270
Milling Practice at Cobalt, Ontario 274
The Nippining Co.'b "Low-grade" Mill, Cabalt, Ontario 274
CHAPTER XXIV Partimq OoLn-eiLvEn Buujon, Pkicbh and Coen:
Parting Gold-ailver Ingota or Ban with Adds 278
Efeetrcrfytic Parting of Gold from Sdver 279
PricM and Coeto 280
PART IV— IRON AND STEEL CHAPTER XXV Ibon Obeb and Thxis SuELTtNo:
Claaeification and Occurrence trf Iron Ores 288
Roasting Iron Ores 286
The Agglomeration of Fine Ores 286
Smelting for Kg Iron 287
Iron Blaet-fumace and Plant 2S7
Gas Cleaning 287
Hot-blaat Stovea 296
Btast-fumaceB and Acoeeaoriea 208
Operation of the Blast-furnace 300
Irregularitiee of Furnace Operation 302
Disiwaal of Pig Iron 304
Dry-air Blast 306
Chemical Reactiona of the Blast-furnace 306
The Heat Balance of the BlastJumaoe 300
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Burdening the Btast-funutoe , 310
General Ammgement of the Blast-funiaoe Plant. . -■ 312
Rf Iron 314
Classification of Pig Iron 314
Influence of the Contained Mementaon the Character of the Pig Iron 316
CHAPTER XXVI Wbouoht Ibon and Steel:
The Manufacture of Wrought Iron by the Puddlii^ Procees 31S
Steel Making 320
Steel Moldng by the Add Beeeemer Proceee 321
Tie Basic Bessemer Procew 326
Steel Making in the Open-hearth Furnace 326
Tlie Open-hearth Rererberatory FXimace 327
TTie Add Open-hearth Proeeaa 33S
Basic Open~hearth Prooese 334
Calculation of Charge 335
lite Open-hearth Building 339
llie thtplex Proeen of Steel Making 340
Duplex and Electric Fumace Plant 341
Electric Steel-making 343
Varietiee of Steel 345
Inn Ore and Pig Iron Pricea 347
PART V— COPPER CHAPTER XXVII Copper Ores and Tbeib Theatmknt:
Characteristice of Copper Dree 3S1
Extraction of Copper from Its Ores 353
CHAPTER XXVIII CoppBR Blast-puknace SifKLTtNo or Oxidiesd OftEa:
Blast-furnace Smelting of Oxidiied Ores 3S5
Lake Superior Copper Country Smdting of Copper Beverberatory Slag 367
Smelting to Black Copper by the Union Miniere du Haut Katanga 359
CHAPTER XXIX Bi/ABT-rusNACE Shxltino of ScLPHinE Ores:
Matte Smelting 860
The Meadter System of Bedding 362
The Copper-matting Blast-f umace 364
Accceaories of the Blast-fumace 367
Blast-furnace Conditions 388
lArge Copper-smelting Blast-furnaces 369
Regular Operation of the Copper Blast-fumace 371
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CoK»r-fum»oe Slags 374
Pyrite Matte Smelting 377
Reactions in Pyrite Matte Smelting 378
Calculation of Charge in Pyrite Smdting 381
Disposal of the Sag 384
Blast-fumaee v». ReTerberstwy Smeltii^ 3S5
Chapter Xxx
The Welsh ProoeM of Reverberatoiy Smdting 387
Smelting Operations by the Welsh Proceas 388
The Direct Process of Rererberatory Smelting 390
large-scale Reverberatory Matte Smiting 390
Tiut Direct Goal-fired Furnace 391
FmiiaoM Kred by Pulverised Coal 392
The Oil-fired Furnaces 395
Operation of a lArge Reverberatory Furnace 397
Seacliona and Calculatifm of the Charge 397
CHAPTER XXXI CoirvEBTiNQ Coppbe-matte:
TTie Copper Converter 400
The Converter Lining 402
Operation of the Basic Converter. 403
Chemical Reactions of the Converter 404
Blast-fumace Smelting and Converting Plant 40B
Electrostatic Recovery of Copper Blast-fumace and Converter Dust 410
Works of the International Smelting Co 410
Coata of a Fropoaed Plant and Operation 414
CHAPTER XXXn
Thh HTDROHvrALLUSQT OP GoPPBit:
Fmeiplea of the Hydrometallurgy of Copper 417
Elxtraction of Copper by Natural or Weathering Methods 418
The Rio Tinto Process 418
The Shannon Copper Company Process 421
Extraction of Copper as a Chloride 422
TIm Henderaon Process 423
Tlie lAist Process 426-
Sulphuric Acid Leaching 429
The ButtfrDuluth Ptooms 430
The Ajo Process 432
Ammonia Leaching ', . . 438
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xiV CONTENTS
CHAPTER XXXIII ItxiTNTNa or BusTEB-coppiiit:
Copper Refining 442
Melting and Refining Uke Copper 444
Tbe Making of Anodes and of Commercial Cathode Copper 446
CHAPTER XXXIV ELBCTBOLmc CoPPKH RspiNiNa:
Electrolytic Copper-rafining Plant 449
Capital Requirements 456
Cost of Refinery and Operating Costs 4S6
Schedule of Copper Ore Prices ifi7
Part Vi— Lead
Chapter Xxxv
Propebtibs or Lead and its Oreh:
Characterietics of Lead Ores.. 461
The Smelting of Lead-bearing Ores 463
SmdtJDg cm the Ore-hearth 463
Chapter Xxxvi
Silver-lead Blaatr-fumace Smelting M7
Receiving, Sampling and Bedding of Lead Ores 467
General Arrangement of a Smelting Works 468
Bedding Ores at a Custom Works 469
Open and Cloeed-tap Bla8t--fumaoea 475
Operating tbe Blaat-fumace 476
Chemical Reactions and Physical Changes of the Blast-furnace 479
Slaga in Silver-lead Smelting 480
Fuel in Silver-lead Smelting 483
Calculation of a Blast-furnace Charge. 484
CHAPTER XXXVII Proddctb or the Bla»t-bubnace :
Flue-duat 487
The Bag-houae 488
Lead-copper Matte 491
Comparison of Matte-treatment Methods 491
Converting of Leady Matte 492
Belling Prioe of Matte 493
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CONTENTS vr
CHAPTER XXXVra
Production op Lead Ombs and Prices:
Ore Prioea; MiMiaaippi VaUej SmeltiDg Wc^u 49S
Variation in Costs Due to OutjHit, etc 4S7
CHAPTER XXXIX RBTiNDra OF iMAB AND Bau Buluon:
Refiniitg Base-buUioD 4S8
TheRefinwy 49B
Softening BaBs-buUkn SOO
The Parkes Piooees 908
Treatment of the Rich Lead 607
The Pattinson Process 610
Cost of Refining Base-buUicin 611
Selling Price of Baas-bullion 511
The Bette Prooets for the EleetnJytic Refining of Lead 511
Smeltery and Refinery for Slver-lead Ores 61S
Part Vii— Zinc
CHAPTER XL Zinc and Its Osns:
Pnqteities of Zine 517
Zine Ores 617
Chapter Xli
RoABTiNo Zinc Ohxb:
Reduction of Ores of Zinc 519
Roasting Blende 51B
Chouistry <rf Roasting Zinc Ores 519
Roasting Furnaces 520
The Wedge Mechanical Blende-Roasting Furnace 621
The Hegeler Fumaoe K21
Various Furnaces 525
TTie Merton Funwcs 626
The Ridge Furnace 526
Sulphuric Acid 627
CHAPTER Xm SifELTiNQ or Zinc Ores:
• ftnelting or Distillation of Roasted Zinc Ores SX
I Operating the Fumaoe 584
Manufacture of Retorts and Condenaera 536
Loss in the Process 537
Cost of ftnelting 638
Price erf fflnc Ows and Bpelt«r in 1919 689
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Contents Chapter Xliii
Zmc RariNiNa;
Grade* erf Spelter. 640
The De S&uUes RedistiUatioii Method 642
BefiniDg Spelter without R«diBtilUtian 642
Electrolytic Zinc 542
PART Vm— PLANT AND EQUIPMENT AND THEIR COSTS
Chapter Xuv
Location, Equii-ment and Erzction:
Location of Works 647
Nature of the Site to be Cbaeen 548
Conatruction of Plant 660
Chapter Xlv
AccBssoRT Equituxnt or Plant:
Intermittent Handling of Materials 651
Industrial Locomotivee 561
Industrial Cars and Hoists 6S3
Grabs and Excavators 666
CHAPTER XLVI Ore Storaob and Scpflt:
Provision tor Supply 563
Feeders 568
Pumps and Elevators 606
CHAPTER XLVII CoBT or Plants:
Coat of Plant 669
Coat of Metallurgioal Plants 571
Unit Construction Coeta in 1914 572
, Composite Coeta 576
PART IX— THE BUSINESS OF METALLURGY CHAPTER XLVIII Tbk Gbnebal EcoNoinc Sttoation:
Distribution of Wealth 679
Economics of Engineering 579
The labor Stuation 580
Financial Crises m the U. 8 583
AaeociatkiD of "Eatpiayea 583
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CONTENTS XVii
Chapter Xux
Organitatioa ot a MetalluTgical Company 5S4
Hjc Adminiatrative Deparbowit 5M
Hw Opemtiug DeportmeDt 685
RukB of Worka 687
Hant Operation SS7
Mwvle of IiiBid»4iien 688
Modes o/ Payment 689
The Accounting DepdrUnent 692
AdminiEtration and General Charges 594
laical Operating Department 594
Ute PuTcbaging and Selling Depart>ment8 695
CHAPTER L Pbofitb and Cone:
PiofitB tm
Custom Smelteriee SOS
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Errata
page 8 — Figure reference 14, 3d line from top, ebould read Figure 4.
page 77 — Figure reference 7, 7tti line from top, should read Figure 14.
iMige 78 — flfrure reference 122, 6th line from bottom, slionld read Figure 212.
page 84 — Figure reference 181A, 7tb line from top, should read Figure 278.
page 93 — Figure reference 31, lltb line from bottom, abould read Figure TO.
page 130— Figure referencca 29 and ,"M, last line should read Figures 60 and «1.
i>age 135 — Figure reference 10, 7th line from bottom, should read Figure 71.
page 2fiO— Figure reference 117, Ist line, should read Figure IIOA.
page 322— Figure reference 104, 12tb line frMn top, should read Figure 18».
I>age 388— Figure reference 29, 16tb line from top, should read Figure 63.
page 40S — fi^gure l!il, should read Figure 221.
page 421— Figure 221, should read Figure 230.
page 446 — Figure reference 147, 1st line, should read Figure 242.
page 478 — Figure reference 156, 14tb Hue from bottom, should read Figure 2(13.
page 523 — Figure 2fZ, should read Figure 282.
page 52S — Figure reference 381, dtb line from top, should read Elgure 284.
Tth Hue from hutton, should read
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Part I General Metallurgy
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ORES AND METALS DKFnirnOH AHD CLASSIPICATIOlf OF ORES
De&iitloiL — An ore from the Btandpoint of the metallurgiBt may be defined as a mineral aggregate containing metal, or metals, in sufficient quantity to make their extraction conunercially profitable. Minerals or rocks ctmtaining 15 to 30 per cent iron would not be called iron ore, nor would we call a rock containing 2 to 3 oz. silver per ton a silver ore. On the other hand, the rock of the Treadwell mine, on Douglas Island, Alaska, carrying (2.50 to S3 in gold, is called a gold ore because it can be worked at a profit. In general, ores are named from their dominant metal (as lead, copper, or silver), though they may contain other metals. Thus a lead ore may contain silver and gold; a copper ore, besides copper, may contwn silver, gold, and even lead. The appearance of an ore may indicate whether it carries lead, copper, iron, or sine, but gold and sUver minerals are not always visible, and the proper way to determine their presence is by assay.
Strai^t or sin^e ores contain in the main but one kind of metal, such as gold, silver, copper, or lead. Straight silver, or free-milling silver ores, are free from lead and copper, and may be treated by amalgamation. Straight gold ores, also free-milling, are those contaiiung the gdd in metallic fonn and amenable to amalgamation. Straight or plain lead, zinc, or copper ores do not contain gold of silver in quantity sufficient to pay to separate the precious metals from the base metal. As an example, a lead ore containing 4 oz. silver per too would not ordinary^ meet the cost of extracting the silver. Blister copper may contain as much as 12 oz. silver per ton and yet not pay the charge for electrolytic refining for its recovery. An ore containiiig little lead, say less than 5 per cent, is designated a dry ore. Such ore is often silicious, but possesses commercial v^ue becatise it contains gold and silver. Ores carrying more than 5 to 10 per cent lead may be profitably treated for their lead alone. Ck)pper ores also frequently contain gold and silver.
Mixed ores, or those containing two or more kinds of metal, are common, such as silver-gold, ailvei^^ld-lead, or lead-zinc-copper-silver. When such ores contain both copper and lead it is puzzling at times to
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4 Ores And Metal£
know how to designate them. In doubtful cases smelting companies have purchased them either on the basis of their lead or copper content under the plea that, in extracting one of these metale, the other is lost or wasted. Lead-silver or lead-silver-goki ores are those which cany lead in such quantity that when the lead is recovered from them by smelting, the precious metals taken up by it can be later easily removed from the lead. Copper^ilver, coppeivsilver-gold, or copper-gold ores, when smelted, yield their copper, and this, like lead, takes up the precious metals.
Bafl«-metal Ores. — Lead and copper ores often contain zinc, antimony, araenic, tellurium, or Insmuth as impurities. These, in the process of reduction, alloy with the principal metal to its commercial detriment, and require expensive after-treatment to remove them. While a free-milling ore permits the extraction <rf most of its gold or silver by simple processes of grinding and amalgamatiwi, a refractoiy or rebellious ore requires pre-limiDsry treatment by roasting before it can be amalgamated; otherwise it must be smelted. Even smelting ores may present difficulties of treatment that would cause them to be called rebellious. A docile ore, on the contrary, is one that may be easily treated. Gold and silver ores containing arenaic or antimony may be cited as examples of refractory ores.
Methods Of Treatubnt
These may be divided broadly into millie^ or smelting plants. Mills treat gold and silver ores, according to their character, by concentratibn (including flotation), amalgamation, chlorination, cyaniding, or by combinations of these methods. Thus, the North Star Mine, Grass Valley, Cal., treats a gold and silver quarts ore (carrying sulphides) by amalgamation, concentration and cyaniding of the concentrates and tailings. The Tonopah-Belmont, a medium hard quartz silver ore carrying sulphides, usee concentration and cyaniding. Another miU, the Liberty Bejl, Telluride, Colo., having a soft quarts edlver ore, subjects it to amalgamation, concentration, and cyaniding of the tailings. Smelting ores may be basic, silicious, dry, coppery, or leady;' while milUng ores may be talcoee, quartzoee, raw, roasting, eartl^r, argillaceous, light, heavy, or base, all of which characteristics modify the mode of treatment. Among the iron ores we may have Bessemer ores or those containing not more than 0.045 per cent phosphorus, and non-Bessemer ores, or those so high in phosphorus that the pig iron made from them needs subsequent treatment in the tmac open-hearth fumaoe to remove it.
An ore consists not only of the species of metallic compound from which it is named, but also of gangue or waste matter. This may often be its principal constituent, and may be earthy, silicious, argillAceous,
Methods Of Treatment 6
talcoee, or limy, aod the ore may be composed largely of the lighter gangue with comparatively small quantities of the valuable metals scattered or disseminated through it. When, as is often the case, the metal ie the heavy part of the ore, and the lighter part is the gangue, the ore m*y be concentrated or dressed with a view to removing this gangue. Ao oie capable of being thus treated is called a concentrating ore, and the valuable heavy part obtained from it is called a " concentrate."
We may also divide ores into sulphide and oxidized. As a matter of fact, these merge into one another, and it is often difficult to decide to which class to assign a given ore. Carbonates are placed among the oxidized ores, since, in smelting, the carbon dioxide is readily driven off, leaving the oxide of the metal.
Grading Ore. — Miners often find it profitable to sort their ore into different grades, such as shipping or smelting, and into milling or concentrating ore, according to the after-treatment they purpose to give it. This matter is often an important one for the metallurgist to consider in deciding uptui the treatment o( ore, as, for example, in the case of a mixed BilTer ore.
Classification Of Hbtaixukgical Operations
These may be roughly divided into two, viz., milling and smelting.
Gold and silver ores are commonly treated in mills, though they may also be smelted.
Iron, lead, sine, and copper ores are commonly smelted, though the last three can be treated by hytlrometallui^cal methods in which the metal is brought into solution and Uttet precipitated from the solution. We speak then <rf smelting beii^ a pyrometalluigical process, while hydro-metallurgy relates to the extraction of the metal by aqueous solutions. In smelting, the ore is roasted if neceasaiy, snelted in a smelting furnace, and the product, the metal still containing impurities, refined to put it in marketable form. Gold and silver have been successfully recovered from free milling ores by crushing and amalgamation. However, most ores cannot be so easily treated. The steps of mJUing practice then are (1) cnishii^ and grinding, (2) solution, (3) filtration, (4) precipitation, (5) refining, as given in later chapters of this book.
FSmCIPLBS RBLATING TO THE REFmiHO OF HBTALS
It ia found by analysis that the separation of a metal from other metals or from contained impmities is seldom complete. It is difficult and commercially impracticable to obtain metals entirely pure, so that those that come on the market still contain small amounts of impurity. Metals thus prepared are graded according to quality, and command pricee
Ores And M£Tai£
Mg Mn Hg
one
Ci
35
Kl
19
H
1
N
14
0
16
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Molding And Casting Metals 7
according to the gmde. Thus Lake copper commands the highest price of any copper because of its purity and toughness, while electrolytic copper adls at one-half cent less per pound.
In alver4ead smelting practice the slag, no matter how thoroughly settled and separated from the matte, still contains 0.2 to 0.3 oz. silver per ton and 0^ to 0.4 per cent lead. In copper refining, in the reverberatory furnace, arsenic, antimony, and bismuth, occurring in the crude or blister copper, are retained as traces after refining, and where t^e bhster copper is impure, no high-grade product can be expected. In the sepantion and deposition d copper by electrolyaiB at low current-denmty, the copper is of hifi^ grade even though impurities are in solution in the dectrolyte; nevertheless, traces of impurity find their way into the cathode copper, though to lees extent than by any other system of refining.
In the refining of pig iron to make steel, in order to obtain satisfactory quality, impurities must be removed until less than 0.10 per cent phosphorus and 0.05 per cent sulphur are present, otherwise the steel lacks toughness and tenacity.
Hoidino And Castiho Metals
Metals undeigoing treatment are finally brought to the metallic state, and are c(»nmoiily cast into ingots or bars for sale. Sometimes metals may be finally granulated, as zinc for cyaniding or lead for test-lead in assaying. Or again, the molten metal may be poured into water, producing coarse flattened granulations where it is desired to quickly dissolve the metal, as in the parting of precious metals.
But generally metfds are cast into bars, ingots or commercial shapes as desired by the cust(»ner, who wishes to subject such bars to further treatment. These shapes vary according to the metal and are cast in molds either by hand or by casting machines as follows:
Gold. — This is cast of all sizes according to the quantity treated, or to the desire of the cust(Kner, from the size of the finger upward, to be rolled into sheets, to be drawn into wire or to be sold to the mint. Gold for the mint is then remelted in plumbago crucibles of suitable site, assayed and paid for, then granulated for parting.
^ver is commonly cast in bars of 1000 oz. (about 80 lb.), a convenient size for handling, and hard to steal. When ready to tap from the test of a cupelling furnace it is run into ingot molds standing on a carriage beneath, then pushed along as the ii^ots fill. In similar quantity Ha metal may be remelted in a plumbago crucible, then lifted out with basket tongs and poured directly into the molds, as in silver-mill practice. Where the quantity of metal is large, as in the treatment of precipitate, Uie melt may be poured into a crucible and thence into molds, as shown
in Fig. 2. The tilting furnace i and pouring position.
oil-fired and is shown in skinuning
¥ia. 1. — Pluisbago Cmciblea.
Fio. 3.— Conical Mold.
Fig. 14 ifi a mold such as is used for silver bars 11 in. long by 4)
in. by 4) in., to hold 1000 oz. or 70 lb. of the metal.
Iron and SteeL — ISg iron from the bla8l>-fumace was formerly cast upon the sand floor of the cast-house. The furnace was placed centrally at one end of the house, and the floor sloped away at even grade from the metal tap-hole. There was a central channel or runway made in the sand to the end of the house. Branching either way at right ai^es from this channels
' were made in the sand, and by
means of wood molds, caviUes for the pigs. Thus the branches were fancifully called the bows, the cavities the pigs. When iron
Fio. 3.— TOting Furnace.
Fio. 4.— Caat-iron Mold.
Fta 5 — Boeh and Molda.
was tapped the flow was down the mam channel To divert the flow, gates (flat cast-iron plates) were so set across the channel as to change
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Molding And Casting Metal£ 9
it to any branch, filling the cavities fonning the pigs. When they were filled the gate waa set at the next branch the main run was opened, and ibs filled side branch cut off, all t^ means of the gates. After cooling the pigs were broken off by prying them up with a bar, and removed to the cars standing upon tracks alongside the cast-house.
The |Hg8 had some saod sticking to them, and this was one reason that mechanical casting was adopted, as in the Heyl and Pattison machine, Fig. 162.
Fig iron, as we know, upon remelting in a cupola furnace, makes the most intricate castings. The same is true of steel castings, now successfully melted in the converter or in. the electric furnace, treated by the addition of ferro-alloys to make a quiet melt, and producing strong "**rf.ingB itx special purposes.
Copper. — This metal after refining is cast into special forms for the market, in small furnaces by hand. The skimmed metal is dipped from the furnace, using a dipping ladle having a bowl 9 in. diameter and holding 25 lb. In casting, a water-bosh is used, see Fig. 5. On the edge of this is hinged a number of molds which are successively filled by three or four men who dip from the furnace and fill them. As soon as these get Bohd, the mold is tutned over, the ii^ot falling into the water. From the water it is picked out by means of tongs and b ready for market. The usual shapes are ingots, ingot bars, wire bars, and cakes, which wei^ respectively 17 lb., 35 lb., 35 to 250 lb., and rectangular cakes 14 in. square or 14 by 17 in., to roll into plates.
In place of the hand ladles, often " bull ladles " are used. These, having long handles, suspended near the center of balance by chain to an overhead trolley rail, will dip up 100 lb. at a time. They greatly expedite the work of dipping. This, with an ordinary refining furnace, will take four or five hours.
In place of hand dippmg, one of the mechanical casting machines, such as the endless mold machine, Fig. 242, or the Walker castii^ machine, Fig. 245, is employed. Indeed, for one of the lai^ furnaces, hand dipping would be too slow. As seen in Fig. 245 the machines are supplied by a ladle from the converter, and blister copper in ingots of 250 to 400 lb. are cast.
In place <^ a casting machine a series of molds are often employed, enough of them to take care of a ladle full of blister copper. When a works desires to produce a finished cathode, this is done by collecting the converted product in a tilting furnace, resembling the tilting open-hearth furnace of steel practice. It is here poled, making a smooth ingot, ot even an anode suitable for use in the electrolytic vat.
I«ad. — The older way cA casting base-bullion was to let the molten metal run into & colder, a basin that would hold 1000 !b. The dross was
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skimmed from this, and the baae-bullion dipped by ladle into jnolds holding 85 to 100 lb., then shipped to a refinery. The present practice is to receive the molten metal into a two-wheeled pot. It is then taken to the droesing kettle, where the dross is removed as described under lead refining. The metal, now free from dross, is shipped away to the refinery. The dross is retmned to the blast furnace. It is a coppery dross still containing lead.
Under head of Refining we give the method used for molding market lead. The lead when solid is removed from the molds by hand. In some cases an endless mold machine is used.
Zinc— This b tapped from a horizontal row of condensers into a ladle suspended by chain blocks to a trolley rail. The collected metal is sldnuned from does and poured into molds. In remelting spelter, a charcoal cover should be used, since hot molten zinc easily droeses. It is for this reason that in brass making the copper is first melted and the zinc added at the last moment before casting.
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Chapter Ii Fdbls
eokeansrtiiiciaJoM. The 2S , "f ""'''"^ Co=l i. a natm^l fuel-altered
from their origu,.i eoS™ k T ™'»'»°'»» in some meMm„ wood through part, hj"^ Xi^° ' °" "'' """"""^ "-i ">»!!» ft™ ;^J»phi.. .t .h;T;l'^^."'XS'»"eoaI,.„th™eiteo,h.TdcoS ^]-d prepared fuel, " and •■ ,J^^ ""^ '"J' >« divided into the KM- The solid (uji, Md charcoal.
ite -
int- 6
ult, I I
by nd
ii«
led Fio- 7. — Table Showmg Geneida of Natural
ice Fueto.
lig recorded geological time, retaining the racite where the volatile conetituenta are *nd immense presBure and finally result in •'efes the work.
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THE HATtlRAL SOLID FUELS
Classification. — A convenient division of the standard types ctf auch fuels may be made into first, those of compact texture, and second, those of woody, fibrous, or earthy tfixture.
A more exact and convenient classification, beginning with the most compact, the highest in " rank," is thus given:
(a) Graphite, native coke, and anthracites, these burnii^ with a non-luminous Same.
(b) Bitumino'us coal or bitumen, biiroing with a luminous flame.
(c) Lignites, peat, and wood, fuels having a woody, fibrous, or earthy structure, burning with a luminous fiame.
The impurities of coal are ash, sulphur, and to a lesser extent nitrogen. Of a ^ven type the standard may be given at 6 per cent ash, and 1 per cent sulphur. In nitrogen 0.75 per cent for anthracite; 1.5 per cent in the intermediate types, decreasii^ to 0.75 per cent in the lignites. These impurities cause a variation in " grade."
The rank of a coal, in changing from peat to graphite (see Table II), shows a progressive elimination of moisture and volatile matter and a corresponding increase in the proportion of fixed carbon and ash. Thus, a typical fresh peat would contain 91 per cent moisture, 6 per cent volatUe matter, 2 per cent fixed carbon, and 0.3 per cent ash. A typical lignite, assumed to have been derived from the peat, contains 43 per cent moisture, 26 per cent volatile matter, 27 per cent fixed carbon, and 25 per cent ash. If frcnn lignite we go on up through the list, we find that while the amount of moisture in the coal steadily decreases the percentage of volatile matter keeps about even with that of the fixed carbon in all the lower-rank coal until the moisture reaches a stable minimum beyond which the percentage of volatile matter is rapidly reduced. Thus we find the ratio of volatile matter plus H20 to the fixed carbon is one-to-one in the lower-rank coals; it has risen to one-to-four in the best bituminous, and to one in seven or more in the anthracite.
The Anthracites. — These have a fuel ratio of one to seven and bum with a non-luminous fiame. They are conveniently grouped into the " hard," having a conchoidal fracture, h^h specific gravity and sub-metallic luster; and the soft with a semi-cubic fracture and low specific gravity.
The Bitumites. — These include the bituminous coals of the carboniferous age and the sub-bituminous coals, or those of the post-carboniferous, both having a fuel ratio less than one to seven and burning with a luminous flame. This flame indicates the presence of hydrocarbons in the volatile constituents, and so that the coal is bituminous.
The coals of the carboniferous age are divided into (1) the BOH»lled smokeless Virginia coals, which includes those having a short and those having a medium flame; (2) the coking or steam coals, having a loi^ flame;
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Composition Op Goals
Is '3
J
• S 3 S ■" S " a S 8 s
III I P 111 II fl
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(3) the non-coking or " household " coals, havmg a woody texture and cubic fracture; household coals also having the same t«xture, but a conchoidal fracture.
The coals of the poet-carboniferous ^e are divided into (1) the weather-resisting or Montana coals, which may be stocked or shipped long distances, and (2) the non-weather-resistii^ or semi-bituminous, which exposed to alternate wetting Eind drying, break down and lose their shape in a month's time. In this latter group is found New Mexico coal, having a fuel-value of more than 7750 calories, and the Wyomii^ coal with less than that.
The lignites are high in moisture, the black lignites or sub-bituminous coals containing much less than 30 per cent, the brown coals more than 30 per cent raobture. They have a woody structure and a luminous flame.
In the following table of coals of the United States, we give the ratio obtained by dividing the fixed carbon, as found in the proximate analyses, by the combined percentages of volatile combustible and moisture. The calorific value of the coal and its analysis, both proximate and ultimate, are also given. Of the two, the former is the one generally thought of. A coal of standard type is that used.
Wood. — When, freshly cut, wood contains 40 per cent moisture, and in this condition is difficult to bum alone; but where this can be done, it develops 2300 pound-calories per pound. Split into cord-wood, piled, and dried for several months, wood contains 20 per cent moisture and 40 per cent carbon. Its calorific value thereby increases to 3600 calories. Such wood is classed as hard when ita specific gravity is more than 0.55; below this it is called soft. While the calorific intensity of dry wood is low, its combustibiUty is great, and it is well suited for use in reverberatory roasting-fumaces, since the volatile constituents, rapidly escaping, bum gradually, and make an extended fiame along the hearth of the furnace, heating it more uniformly than could a Sameless fuel like anthracite or coke. In the outlying districts of the western United States, where the metallui^t is dependent on wood for generating steam, or roasting ore, the accumulation of a sufficient supply of dry wood should be one of his first cares. In this his forethought is well rewarded. He should purchase wood delivered and corded near the works; and in measuring, make equitable allowance for short dimensions or open piling. Ck>rdwood should " cord up " to 70 per cent solid wood.
Coal for Roastiiig.— Both lignites and the regukr bituminous coal may be used. It should have a good proportion of " volatile matter," so that the flame may be long, and thus distribute itself over a greater area of the roasting hearth. A short flame would be intense near the fire-box, but would fail farther away. To make the flame long the quantity of air is 30 regulated that it mingles slowly with the escaping volatile gases, and thus it is carried to the end of the hearth.
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FUEL Oils ;i5
Coals of very different properties may appear alike if represented only by proximate analysis. The comparative calorific value may be judged of by Berthier's method. This consists practically in the operations of a lead assay, tising an excess of litharge, with a gram of the fuel, and noting the size of lead button reduced. One can also judge a good deal about the character of the coal by coking it in a covered crucible and weighing the coke produced, juicing the character by the appearance of the product. The proximate analyses (Table II), showing the different kinds of coal, determine to which class any given kind below^.
Gra^iite. — This is of interest, not as a fud, but as a refractory material, particularly when combined with clay.
Petroleum or Fuel Oil. — This is the most ccmcentrated of fuels, and, when the coet justifies, can be used not only for generating steam, but for roasting and melting. It will be found, in bumii^ fuel-oil from various loc:ilities, that the calorific power is much the same for the different kinds. Beaumont (Texas) oil tias a calorific power of 10,820 calories, and a specific gravity of 0.88 (7^ lb. per gallon.) Oil con be burned in sucb a way as to give, not only a high and uniform temperature, but also the oxidizing (roasting) or reducing action that may be desired. The air for combustion is best preheated as well as the oil, and it will be found advan-ti^;eou8 to inject the oils under a high steam pressure. A mixture of light and heavy oils should not be used. In Russia, where it has been employed in open-heartii steel-furnaces of 10 to 16 tons capacity, oil to the extent of 15 to 20 per cent of the weight of the chai^ has been used. As r^ards comparative costs at the Selby Smelting & Lead Works, Vallejo Junction, CaUfomia, it was found that the saving was 40 to 60 per cent with oil at S1.71 per bbl. (42 gal.) and coal at S6 per ton. A suitable control df the grade of the matte was possible by the r^^ation of the fiame.
Natural Gas. — In Ohio, Indiana, and Kansas, particularly, there are districte where natural gas has been obtained by boring for it as for oil. It is the most efiicient of natural fuels, having a caloriSe power of 611 caJ. per cubic ft. or 27,862 Cal. per pound. The following analysis will give au idea of the composition of Pennsylvimia natural gas. It shows that it is composed chiefly of marsh-gas and hydrogen, vis:
Per Cent. by Volume.
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THE ASTinClAL FUBLS
These include charcoal, coke and producer gaa, aJl made from natural fuels. ,
ClurcoaL — Wood, packed in a kiln, and pennitted to partly bum, changes into charcoal by distillation of the volatile portion by the heat produced from the portion burned. The charcoal retains the form of the wood from which it was made, but has a specific gravity of only 0.2. It is of a dull-black color, soils the fingers but slightly if of good quality, but much if poor. It should ring when struck, and should show the annual rings of the wood distinctly. The density of charcoal varies with that of the wood from which it was made, dense woods giving a dense charcoal. A heaped bushel (1.5555 cu. ft.) weighs 14 to 16 lb. When apparently
Fig, 8. — Section of Charcoal Kiln,
quite dry, charcoal still contains 10 per cent or more of moisture. Dry charcoal contains 95 per cent carbon, 1.5 ash, and baa a calorific power of 7610 pound-calories per pound. Charcoal is used in iron blast-furnaces, particularly in locahties where wood is abundant; and it produces a pure, strong iron, free from sulphur, called " charcoal-iron." Charcoal has been used also for silver-lead and copper smelting in districts difhcult of access. In these cases it has done especially well when coke could be secured to use in conjunction with it. It is, however, a friable fuel, making fine dust sometimes to the extent of 10 per cent; and this " fine " is apt to make trouble in the blast-furnace. If under-burned, it is heavier and more dense, and has a brown color. Portions of the wood found imperfectly burned are called " brands " and are returned for the next burning.
Charcoal is generally made in a kiln. One of these in section in F^. 8, shows the method of filling. The kiln is set at the foot of a steep bank, so that it can be chained conveniently from above. It has two charge-doors A and B. The first of the wood is conveyed through the lower door,
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Charcoal 17
and placed. The remainder is brought along the nmway C, and introduced through the upper door B. There are three rows of openit^, 3 by 4 in. in size, spaced 2 ft. apart, around the bottom of the kiln. The kiln is lifted at the lower door, and when fairly started, both openings A and B are closed with sheet-iron doors. These are tightly luted with clay, and the air is thus cauBed to enter by the small holes. When combustion has progressed sufficiently, these openings are ti^tty closed, and the kiln is permitted to oofrf slowly. The period of charring or burning is eight days and the cooling four days additional. Such a kiln holds 25 cords of wood and produces 1125 bu. of charcoal wei^iing 16 lb. per bushel, or about 20 per cent of the weight of wood charged.
By-product Charcoal.— An example of the modem method of makii^ by-product charcoal for iron blaet-fumace use ia one at the Pioneer Iron furnace, Marquette, Mich. Here there are 86 kilns each holding 8 cords. The daily requirement is 20 carloads, of 16 cords each, amounting to 320 cords. The kiln is packed full of wood, the sheet-iron doors put on and closed, and fire is started at a manhole in the apex of the dome. As soon as combustion gains sufficient headway, this opening is closed, and smoke escapes by way of a flue leading from the base of the kiln to the chimney, continuing thus until most of the aqueous vapor has escaped. At this stage the chimney is closed, and the vapors pass by a smoke-main to the condensers, the current being aided by an electrically driven fan. The cold Biu^ace of the copper tubes of this condenser precipitates the condensible portion of the gas, while the gas itself goes on to the boilers, where it is burned for steam-making. The condensible portion, amounting to 41 per cent of the weight of the wood, is called green liquor or pyroligneous acid, and consists mostly of water, but contains also alcohol, tar, ammonia compounds, acetone, and acetic acid. The tar ia separated in settling tanks, and the liquor passes to the primary still-house. Copper stills here remove the vapors of alcohol, acetic acid, and much water frran the liquor. The neutralizing tank receives the product, and into this is mechanically stirred milk-of-lime to neutralize the acid by the formation of acetate of lime. The neutralized liquor is allowed to settle, and the supernatant solution is dmwn off and conveyed to the refining-etill house. By fractional distillation a crude wood alcohol is obtained here, and a solution of acetate of lime is left behind, and recovered by evaporating the solution. The crude alcohol is then purified by further distillation until a clear 95 per cent wood alcohol is obtained. A cord of wood (4500 lb.), yields 880 lb., or 19.5 bush, of charcoal of 20 lb. per bushel, 208 gal. of pyroligneous acid, 8 gal. of wood-tar, 64 lb. gray acetate of lime, and 4 gal. wood alcohol. By the sale of the wood alcohol, acetate of lime and formaldehyde, and by the superior quality and consequently higher price of charcoal-iron, it has been possible to build up
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Ftjeis
this industry, where the supply of wood is abundant, in spite of the serious-competition of iron smelted in biasM umaces using coke.
Coke. — This is made from coal in kilns, in a way similar to that of making charcoal. Bituminous coal which cokes or fuses at the high temperature of the kiln or oven is used for this purpose.
The raw screenings, in the example below, contained much fine passing a li-in. bar-6creen. From this, the residue left after removii^ the lumps of merchantable coal, coke was made. By "washing," the fixed carbon was increased and the ash in the coke reduced to 14.24 per cent, A part of the sulphur also was removed thereby. The refuse was high in ash, and low in fixed carbon, as was to be expected; but the yield of washed coal was 85 per cent of the raw screenings, and the coke 70 per cent of the washed coal. When the coal contains slate, " bone," or pyrite, it is improved by this process of washing, or separating the waste-matter by concentrating. An example of a semi-bituminous southwestern coal is shown below:
Mobtiirs.
Volatile
A,h.
Sulphur.
Composition of Coke. — The ash in coke varies from 10 per cent to 22 per cent and the fixed carbon from 77 per cent to 89 per cent- In coke, high in ash, not only has the ash to be smelted, but the fixed carbon is correspondingly low, so that such coke is less efficient. A great difficulty with high-ash coke is that it is often friable, making accretions or scaffolds in the shaft of the blast-furnace. Analyses of two typical samples of beehive coke give the following:
Connellsville Coke: fixed carbon, 87.5 per cent; ash, 11.3 per cent; sulphur, 0.7 per cent. El Moro coke: fixed carbon, 77 per cent; ash, 22 per cent (when the coke, as in this case, is made from unwashed coal); sulphur, 0,9 per cent.
The Coke Ash.— In computing a furnace chai^, this is taken into account. Asb of Connellsville coke contains SiOs, 44,6 per cent; Fe, 15,9 per cent; CaO, 7 per cent; MgO, 1.9 per cent. Ash of El Moro coke has Si02, 84.5 per cent and Fe 5 per cent. It will be seen that this latter ash has a large excess of sihca to be fluxed, and is accordingly less desirable. On the basis of 11,3 per cent ash in the Connellsville coke we would have Si02, 5.0 per cent; Fe, 1.8 per cent; CaO and MgO, J per cent of its total weight.
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Coke 19
Beehive Coke. — A beehive-oven (see Rg. 9 at B), is charged through a hole io the roof. Each oven holds 5 to 6 short tons of coal. In Pennsylvania an oven yields per week two charges of 48-hoiir coke and one of 72-hour. The charge in makit^ 72-hour ooke is dropped in the morning into the hot oven from a coal larry or car above, and is leveled through the side door, filling the oven to the depth of 26 in. The door is then walled up with dry brick and plastered over, but an opening is left near the top, as shown in section, for the admission erf air. Combustion from the red-hot brickwork soon begins, and a dark smoke eacapes at the top opening. After four hours this becomes dense and white, and the gases ignite or strike, and flames issue from the top. For twelve hours the oven bums with a dull, smoky flame above the surface of the charge. The flame be-
Pia. 0. — Sectbns of By-product (A) aod Bee-hive Coke Ovena.
comes br^jht by the second day and then the air<supply is partly cut off. On the third day still less air is admitted, and at the end of this day no more flames appear and the whole interior of the oven is red-hot. The air-openings are now luted, and the chaise is left in this condition until the morning of the fourth day when the coke is drawn. The actual coking is complete in fifty-five hours, and the whole operation, from one charging to the next, in seventy-two hours. To draw the coke the temporary brick wall of the door is taken down, and water from a hose played into the oven. After being thus cooled on the surface, the coke is pulled out with a )ong-handled coke-drag or hook, and further cooled with water while being withdrawn.
The process of fusing and coking begins at the top, and extends downward throu^ the mass of coal to the bottom of the oven, and the coke, when well burned, takes the form of prismatic masses, see Fig. 9, with
hard side-surfaces of a silvery steel-gray color, and top ends soft and nearly black. The silvery appearance is due to deposited carbon, which has the desirable quality c^ protecting the coke against the acti<Hi tA the fiimace-gases. The black ends, on the contrary, are readily attacked. A good coke has a well-developed cell structure which permits ti»e penetration of the hot ascending gases in the blast-furnace. This so raises the temperature of the coke that the air, at the bottom of a furnace striking it, produces vigorous and rapid combustion. Other qualities are purity, unifonn quahty, and sufficient coherence for handling. Purity depends upon a low ash, 10 per cent being good, and 6 or S exceptionally pure. Coke intended for iron blaBt-futnace work should hot contain more than 1 per cent sulphur and commonly leas than 0.5 to 0.8 per cent of this dement. For lead or copper blaet-fumaces high sulphur does not greatly matter. " Uniform quality " means but a small amount of " black ends." These as stated, bum in the upper part of the iron blast-furnace by the action of carbon-^nonoxide gaa. " Coherence in handling " as is evident, is important where coke must be transported far, and rehandled at the smelting works. lines tend to " dow down " a blast-furnace, but can be rejected by the use of a coke-fork. The calorific value of Pittsburg coke, containing 89 per cent fixed-carbon, 10 per cent aah, and 1 per cent sulphur, is 7272 Ib.-cal. per pound.
The By-product Coke Oven. — This Is coming increasingly into use, due to the fact that in it, the products, aside from the coke, can be saved and sold to advantage, and not wasted as is done in maldi^ beehive coke. Moreover, due to the constant h^h and quick heat produced, it can coke a coal that contains but little fusible matter or is nearly non-coking.
A beehive oven will yield 2000 lb. of coke from 3200 lb. of coal, a by-product oven from the same amount of coal will yield 2300 lb. coke; 19 gal. tar; 42 lb. ammonium sulphate; 4.5 gal. benzol (motor fuel) and 10,000 cu. ft. of fuel gas.
The types of ovens of this kind extensively in use are the Otto-HofTman, the Semet-Solvay, and the Koppets. The longitudinal section of a Semetr^lvay oven is shown in Fig. 10. The croaa-eection throi^h the ovens themselves is given in Fig. 9.
Coal is brought in over two tracks, and discharged into feed-hoppers. It is drawn from these as required, and conveyed to two seta of rolls, one for coarse, the other for fine crushing, and reduced to a size of 4 to 10 mesh. The crushed coal is raised by an inclined elevator, and discharged into the xrwa storage coal-bin. This bin has a hopper-ehaped bottom with several dischatge spouts, deUvering to an 8-ton larry which runs along on top of the ovens or retorts, of which there may be 20 to 60, placed side by side, in <me block of masonry. Each retort or coking chamber is 17 in. wide, 43 ft. 6 in. long, and 6 ft. 6 in. high, and is closed at each end by an air-tight
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By-Product Ovens 21
cast-iron door. In ¥^. 9 is ahown a transverse section of such a chamber with the interesting lines of fractures and columnar structure of the coke indicated.
At Fig. 10, the charge-car is seen above the retorts. It is worked by an electric motor and consists of 4 hoppers supported b^ a frame upon a traveUng carriage. The doors of the chamber being closed, and the chamber itself hot from previous operation, a chai^ of 8 tons of coal is dropped in, and leveled by means of the top bar of the chargmg machine, Fig. 12, inseriied through an opening near the top of the door. Distillation
Fio. 10, — Semet-Solvay By-product Oven.
at once begins, and the gases are conducted to condensing-chambers to free them from certain by-products, such as tar, ammonia, and benzol. The first portion of the gas is highest in illuminating power, say 24 candle-power, but later drops to 16 candle-power. The first is, therefore, sent to the city-mains for use as illuminating gas, the latter reserved to heat the chambers by combustion in flues which encircle them. These flues are beneath the chambers, and the side-walls are constructed to provide them for hpiftfing the oven, and maintaining the activity of the distillation. The products of combustion, before entering the stack, go throi^ a regenerating chamber containing a checker-work of tile, while air is preheated for combustion-in a similar chamber at the other side. Thus the gas is burned with highly heated air, and produces an intense heat in the walls of the coking-chambers. The reversing valves are now changed, and the currents of air and gas caused to move in the opposite direction. The
Ro. 11. — Koppere By-product Coke-ov«i.
Fio. 12.— Perspective View, Ovens, Coke-pusher and Leveler.
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Gas Producers 23
direction is thus repeatedly alteraated, as is customary in open-hearth work. At the end of twenty-four houra, when coking is complete, the end doors are opened and the coke is pushed out by means of a coke pusher, Fig. 12. The pusher-head is shown at the left in Fig. 10. The coke is received in coke car, shown at the right of the oven, and is here cooled with water. The total yield of coke is 72 per cent, or 6 per cent more, for the same coaJ, than that of a beehive oven. The coke is hard, dense, and as reliable as beehive coke made from the same coal, but has not the silvery ^oss of - the latter.
Costs. — The actual cost of making coke may be stated as 50 cents per ton in the beehive process and 37 cents in by-product ovens. To this must be added the cost of the 1) tons of coal required. A beehive plant operated six days per week and of 400-tOQ daily capacity would cost S60,000. A by-product plant of the same capacity would cost $300,000. Allowing for interest and depreciation, the cost is found to be much the same for either process.
Producer-Gas
Of the various kinds of producers used for making artificial fuel-^as we shall consider two, " the simple producer " and the " mixed-gas producer."
Fia, 13. — Section View of Gas-producer (band poked).
The Simple Producer. — These use ordinary or inferior fuels, such as rood, wood-refuse, bark, sawdust, or peat, but generally soft or hard loal. We have shown in Figs. 176 and 177, in the sections ot furnaces
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containing fuel, how gas ia produced where air rises through a deep coke fire and where fuel is thus in excess.
Fig. 13 is a simple producer, the neceeeary air being suptdled by a natural draft or by a fan. The fuel, descending in the producer, first is dried by the hot, risii^ gases, then further heated until the volatile matter is distilled, and finally, as it'reaches the lowest zone, is oxidized or burned by the enterii^ air. The residue is the ash of the fuel, which is withdrawn at the bottom. The escaping gases issue at a temperature of 300° to 1000° C.
In the operation of a hand-poked producer, the coal, when charged, if left to itself would soon bum, leaving holes in the fuel-bed, through which would come up unconsumed air. To avoid this a long bar is run down throi^h one of the poke>holes shown in the producer top to breEdc up the hung-up coal and again make the bed continuous.
An analysis of producer^as made from soft coal gave the following resulte by volume:
lOO.O Each pound of coal will give 60 cu. ft. of such gas, having a heatii^ value of 82 Cal. per cubic foot.
The Mixed-gas Producer. — This is the producer commonly used. In it some steam or water vapor is blown with the air into the burning fuel, and t^re reacts upon the carbon as follows :
H20+C = C0+H3
One volume of steam makes one volume of carbon monoxide and one of hydrogen. The steam may be obtained from the water-soaked ashes by evaporation in the lower part of the producer, or as in Fig. 13 may be injected under pressure into the fire. Steam also disintegrates the clinks and facilitates its removal. The carbon dioxide formed means the production of heat later absorbed in the formation of water^as and hence is ui a way useful, since the formation of water^^^ can be carried farther.
The Hu^es Mechanically Poked Continuous Gas-producer. — Pig. 14 is a sectional elevation of a plant containing a row of Hughes producers
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Gas Producebs 25
which give a mixed gas to supply open-hearth fumaoea used in making steel.
The special feature of this type of producer is the mechanical poker F, which is a water-cooled steel casting suspended and secured to a shaft S. The poker is actuated by a mechanism which moves the poker back and forth, agitatii^ and breaking up the mass of fuel, in the slowly rotating shell, evenly distributing the coal, and hdping to work the ashee down-
Fia. 14. — Hn^iw MechuucaJly Poked Continuous Qas-producer Pknt.
ward. Thus the labor of hand-poking is eliminated and the fuel is regularly stirred. This uniform treatment has proved of great advantage, giving uniform results in quahty, quantity, and supply of gas, with a reduction of operating costs.
The brick-Jined producer shell G, or body, is of steel having a cast-iron bas»-ring and a cast-iron water-e(»led ash-pan J and a turn-table R, ill bolted together in one. The turn-table fi has at its out«r circumfer- EDce a cast-iron rack into which meshes a spur pinion keyed to the vertical
shaft S and connected to the horizontal main shaft hy a train of gearing. Thus the turn-table is rotated and with it the ash-pan, base-ring, and body G. The bottom of the turn-table is fitted with a steel tied resting on six conical chilled-iron carrying-wheels. As the producer slowly revolves, the
Fio. 15.^ — Cross-Hection of Revolving lircentrie Ga!4-|jroducer.
ashes work down into the water-eealed ash-pan from which they are shoveled directly into a car in order to remove them. The steel producer top is secured rigidly to the floor structure and a water seal is formed by a flange at the outer circumference of the producer cover. It carries the poker mechanism, two charge-hoppers P, and the gas outlet or off-take, 7*.
0A8 Producers 27
The tatter leads to a gas main or flue connected to the whole row of producers. The base of the producer has a blast inlet-pipe H with a cast-iron deflecting plate for covering the air opening. Air is dehvered to this pipe by means of a blower. The producer is generally driven Iq' an electric motor requiring three electric horse-power.
A hand-poked producer has a capacity <^ 10 lb. of coal per square foot of grate area per hour, while the mechanically poked producer can burn on an average 25 lb. This flgures out approximately one ton of coal per hour for the Hughes producer of 10 ft. internal diameter. An average quality of gas can be maintained of a raimpOKition as follows:
CO, 00 Hydrocarbona H N
4 per cent 22p<'rc«Dt 3 lo 4 per cent 13 per cent 53 per cent
Fio, 16. — Loomis-Pcttibone Gas-making Plant.
Revolving Eccentric Gas Producer. — Fig. 15 illustrates a plant containing a double row of gas producers having grates eccentrically set so that as they revolve they carry the fuel reciprocally to and from the interior walls of the producer shell. The hoppers are filled, and the charge dropped as in the bell of an iron blaat-furnace. By the revolution of the grate the ashes work to and under the peripheral edge of the producer shell. Formerly needing 1.75 tons of coal to give gas enough to smelt 1 ton of charge in the open-hearth, the producer can now do the same work with a consumption of but 0.7 to 0.9 ton.
The Loamis-Pettibone Gas Apparatus. — Fig. 16 shows a complete plant of the Loomis-Pettibone system, with a positive gas exhauster. It is intended both for producer-and water^as. Its operation is as follows: Hot liree are burning in both producers or generators, and the gas exhauster
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is in operation. Air is now drawn upward through generaior 1, burning the fuel and making producer-gas. This generator may have just received fresh coal at E, and the coal-smoke, tany matter, and producer-gas from it, are together drawn down through the hot fire in generator 2, being completely burned and fixed Id so doing. The gas now goes through valve B to the boiler {val*e A being closed), and the heat is there absorbed. It then passes from the top of the boiler through the pipe shown to the bottom of the " scrubber," a cylindrical tower of sheet-eteel, in which it is caused to pass upward through pieces of coke resting upon perforated trays. The coke here is kept wet by means of a water-spray, and the gas is thereby cooled and cleaned. The water drains off by the water-sealed pipe V. RiBing to the top and to the wider part of the tower, the gas passes through a layer of fine shavings or " excelsior," to remove any remaining dust. It is then drawn through the Root positive-blast exhauster W, and finally is driven through pipe Z to the gasometer for producer-gas, where it is stored for use. The fire in generator 1 having become clear and hot, generaior 2 is charged afresh, and the ash-pit door opened. The gas current is then chaiged from generator 2 to generaior 1, through valve A (valve B having been shut) to the boiler, thence through the scrubber and exhauster W, to the gasometer. The direction of the current is thus changed at intervals. For makit^ watef'^as, the ash-pit door is closed and steam from the boiler is injected beneath the grate of the generator while the fire is hot. The formation of the water-gas is completed, or the gas is " fixed " l^ causing it to pass down through the other generator, it having been found that a part of the hydrogen reverte to steam without so doing. The making of water'^as cools the fire and after a few minutes the steam must be shut oS and air again substituted. While water-gas is being made, it may go to the gasometer through the pipe Z, or, if desired, to permit it to go to the water-gas holder, Z may be closed and Y opened. When kept separate, water-gas is reserved for certain beating operations for which producer-gas, of lower calorific power, would be unsuited. The purge-pipe X is opened when starting, and by this means an-in the system is expelled before gas is turned into the gasometer. Steam may also be admitted above the fire, and thus caused to pass down through the generator and form water^as. In fact, both air and steam may be introduced, either below or above the fires, to suit the best conditions of operating.
Comparing the two systems, the band-poked producer costs 70 cents, whi]e the Hughes producer can be operated for 50 cents per ton of coal burned. At the same time a Hughes mechanically poked producer installation is estimated to cost $38,960 as against (45,200 for a hand-operated
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Pulverized Coal
Folverizbd Coal
This is prepared from run-of-mine or from slack coal. It is delivered by car into storage bins, whence it is drawn off upon a conveying belt to the feed hopper of a dryer, Fig. 72. The head pulley at the delivery end of the belt is magnetized, so that stray pieces of iron and steel are removed. In this way trouble is avoided in the grinding machines later on.
Dried down to 1 per cent moisture the dryer discharge feeds to a slightly corrugated roll set to reduce it to pea size. As fast as crushed this product is raised by a belt elevator to a stock-bin to be drawn off as desired to the Raymond roller mill, where it is to be pulverized.
Fia. 17.— Raymond Roller Mill Fia. 18.— Raymond Roller Mill.
Fig. 17 is a view of the roller mill, where at the left is the spout from the stock-t»n, having at its foot a deeply corrugated feed roller called " star feed." This, as it revolves, gives a regulated supply within the truncated casing, where the grinding is performed. As in the Huntington mill, there is a " bull ring," between which and the three suspended rolls the grinding is performed. A head on the central vertical shaft has three suspended shafts carrying the rolls, so that in rapid rotation they are BtroDf^y pressed against the inside of the bull ring, quickly pulverizing the coal. The suction pan above draws up the pulverized material, deUvering it tangentially to the lai^ collecting cone at the right, where it is whirled In cyclone fashion to the periphery to settle to the point of the cone. The air, thus freed from the bulk of its contained dust, is returned by the pipe nsing from the top of the cone to the lower large exterior casing, and passes
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upward inside the bull ring. The coarse particles, th&t have escaped grioding, and fallen to the bottom arc lifted by plows that throw them up, and with the aid of the upward wind current, again lift them to the grinding zone.
The pulverized coal from the roller mill is taken by a screw conveyer to a pulverized coal-bin of 25 tons capacity. Thence it goes by other' conveyors to the respective roasters. Tlie coal must be so finely powdered-that 85 per cent of it is of minus 200 mesh, since the finer it is ground, the greater its efficiency.
The method of feeding powdered coal to a reverberatory furnace is described under head of "pulveriaed coal-firing."
Tbe Hcdback Powdered Cool Distributing System. — This comprises an air-supply pipe into which, under fan pressure powdered coal is fed by feed-screws from coal-storage bins branching to supply all the furnaces. A return pipe takes the excess of air, together with the unused coal back to the storage bins. As more air is used so the coal supply is automatically increased at the feed screws in direct proportion. A control valve is provided at each branch as also a burner for the coal and air to each furnace.
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Chapter Iii
Refractories
Sbfractort Materials And Their F«Oi>Brtibs
General. — The foundations of a furnace may be of concrete or of stone ■aid in lime-mortar, the moderately heated exterior of common building brick also laid in lime-mortar, but for the interior lining it is neceseary to use refractory material to withstand the high temperature and to rasiHt the scouring and corroding action of the molten content* of the furnace. At a temperature below a red heat the combined moisture of lime-mortar would be expelled, and the mortar in consequence would crumble. At a dull red heat many stones crack and flake ofE at the surface of irregular expansion. Sandstone, however, is resistant to fire, and has been used for furnace lining. Red bricks, laid in clay mortAr, withstand a moderate red heat, but, at a temperature much above this, begin to soften or melt.
Refractories.— These substances are infusible at the high temperatures for which they are intended. Thus firebrick only begins to soften at 1500° to 1600° C, and silica brick at 1600° to 1700° C. Refractories may be divided into the three following classes:
Add (Silica-brick, Sand and Ganister). — These are used to resist the scouring or corrosive action of acid slags. Being highly refractory they are more generally used for roofs or arches exposed to the highest temperatures. In such positions out of contact with the molten contents of furnaces they are not required to resist a serious fluxing action.
(2) Heutral (Graphite, Chrome-iron, Fireclay, Bone-ash, and Carbon-brick).— These materials well resist the action of neutral slags which are □either basic nor acid. In the case of a basic open-hearth furnace, for example, it is customary to interpose a layer of neutral chrome-iron brick between the roof cf silica-brick and the basic-lined hearth slightly above the level of the surface of the molten contents of the furnace where it would be unaffected by it. Were silica-brick used in contact with the basic lining, they would react upon the lining and melt.
(3) Basic (Dolomite, Bfagnesite, etc.) — These are used wtiere the slag or matte is basic, as in the hearth of the basic open-hearth furnace. Basic slags quickly scour or corrode an acid, or even a neutral lining. It will be !K>ticed that aU the foregoing refractories not only have special resistant
power but are infusible. This is particularly the case with carbon, either in the form of gas-carbon or charcoal.
Acid Bxpsactoribs
Sand. — This is used in repairing or fettling the interior borders or walls of reverberatory furnaces. It is made to form a steep bank extending above the level of the molten bath, to protect the wall from the corrosive action of the molten slag. Repairs are made after the charge has been withdrawn, when the interior sides of the furnace are expc»ed. In copper reverberatory work the sand is thrown in by means of shovels, or placed by paddles or spoons provided with 16-ft. handles to allow the sand to be dropped at the exact spot required. Sometimes a Uttle clay is incorporated with the sand that it may be formed into balls. These are skillfully thrown across the furnace through a door to an eroded spot, or inserted by means of the paddle, mentioned above, and pressed into position with the bowl of a long-handled ladle. The bottoms of reverberatory fumacesare frequently made of sand in layers, and each layer fired upon and melted successively, at the highest temperature of the furnace. The sand, fritted together, and hardened into a coherent bed in this way, is built to the thickness of perhaps 2 ft.
Ganister. — This is used for furnace-or converter-lining in copper work. It is composed of a mixture of crushed silicious rock or quartz to which has been added about 15 per cent clayey material to make it cohere. For acid-lined copper converters, a silicious ore carrying gold and silver may be used instead of barren quartz rock. The material is rapidly eaten or scoured away by the action of the molten charge, and the precious metal contained enters the charge. This in reality results in a kind of ore-smeltii^, performed incidentally, and without additional cost.
^lica Brick. — When quartz or sandstone, containing 98 per cent sUica, is moistened and mixed in a wet pan (Pig. 170) with a little lime paste made from quick-lime, it coheres sufficiently to be molded into brick. These are first dried in a steam-heated drying-room, then carefully placed in kilns in open order, and burned at a temperature gradually increasing to a white heat. Fig. 19 represents a kiln of the down-draft type. It is a dome-shaped oven, 18 to 30 ft. diameter, coal-fired by means of fireplaces set in the exterior wall. The flues within this wall are arranged as shown, BO that the entering flame rises to the crown of the arch, and, passii^ dowikward through the brick, goes to the adjoining stack through flues in the floor of the kiln. Thus a high even temperature is obtained, and the brick becomes sufficiently sintered to stand handling and transportation, though never as strong as the fireclay brick.
Besides the lime-bocd brick, above described, made by the addition of
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Brick Making
lime to silica, a clay-bond brick, lees refractory, ia made by the admixture of four parts of flint with one of clay. This makes a stroi^er brick thaJi the lime-bond. The composition of each of these kinds of brick is aa follows:
Li IDE-bond
—Brick Knn.
The clay-bond brick shows its greater fusibiUty in its alumina and silica ratio, as will be seen under the constitution of firebrick, and tho pro[>ortioQ of alkali is higher than in the lime-bond brick, causing it to be much less refractory Silica brick withstands the highest temperatures, and expands when heated To provide for this, expansion joints are arranged in the roof, side walls, and bridge of reverberatory furnaces, which close as the temperature rises. To slack off the tie-rods, also, ia another way to accomplish the same purpose. Without this, furnace arches would bulge, and tie-rods would break. The linear expansion of these bricks when elevated in temperature to a white heat is 2.5 per cent.
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Txeotral Ash Basic Refractories
Gr^thite or Plumbago. — Pure carbon in the absence of air is pennanent and infusible at the highest temperatures. This is well exemplified in the carbon filament of an incandescent lamp. Even in the arc4ight, the carbons, thoQgh gradually consumed, do not melt. In blast-furnaces, pulverous carbcm accumulates and forms scaffolds, and carbon-brick, made of gas-carbon, has been used with some degree of success for the bosh-lining of iron blast-furnaces. Graphite is essentially carbon, but contains as impurities a little iron and a small quantity ot gangue substance. An analysis of Canadian graphite gives 2 per cent volatile matter, 20 per cent ash, and 80 per cent carbon. Such graphite is used for graphite or plumbago crucibles and retorts, when mixed with 45 per cent air-dried day and 5 per cent sand. Graphite in these mixtures is not only refractory, but prevents shrinking and cracking when the crucible or other object is dried after being formed.
ChrtHuite or Chrome-inm. — This is a double oxide of iron and chromium (FeOCrjOs) generally containing a little gai^ue. Chrome ore is made into bricks by cruBhing the ore, mixing with lime as in making silica brick, and burning. These bricks should not contain more than 40 per cent CrsOs. Chromite is not attacked by silicious slags, and resists high ' temperatures.
Fireclay, Firebrick, and Tile. — These refractories are the best known and the most used. The term fireclay applies to kinds of clay capable of withstanding a high degree of heat. In good hreclay the total percentage of fluxing impurities, such as ferric oxide, lime, magnesia, and the alkalis, is small (3.5 per cent or less). In all fireclay the water and some of the silica is combined chemically witli the alumina. Thb forms a hydrous aluminum silicate, called kaolinite. Further silica present Is in the form of quartz sand. Either kaolinite, or quartz alone, has a high fusion point (1850" C), but in mixture, the fusion point is lower, and this reaches a minimum at 1670° when 10 per cent kaolinite is present. By the continued addition of silica to kaolinite we therefore get a diminution of refractoriness until this exact proportion is reached, and after this, by continued addition of sand, an increase. The fireclay, accordii^fly, is most refractory that contains the lowest percentage of fluxing base, and the least uncombined sand. A factor further affecting the refractoriness is the coarseness of grain. The New Jersey air-dried clays have the following composition and refractory qualities:
Per Cent. P« Cent.
Free silica 40.09 0.24
Total filming baaca 2.53 0.99
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Brick Making 35
Pv C«nt. Pa Ceot
»0i OT.26 46.75
AlA 23-36 39.06
100.09 100.26 Temperature of fusion 1670' C;, 1810' C.
The clay base is computed as AljOs, 2Si02 with combined water. The silica not present in this combined form is regarded as " free." It is se^i that the leas refractory clay (IV) contains more Suxing base, more silica and less alumina than (V) to account for ita fusibility. The first (IV), ia harder than (V) because more fusible, and is an acid brick, whereas (V) ia neutral. The second (IV), is a type of most of the Western firebrick.
Fireclays are used not only for firebrick and tile, but also for muffles, retorts, and day veBsela of different sorts. The clay varies much in plasticity. Clay alone is unsuited for brick, since in burning it Bbrinks and cracks. Firebrick manufacturer, therefore, employ a mixtiu^ of one or more grades of clay, adding also a certain percentage of coarsely ground firebrick called " chamotte." The addition of this uushrinkii^ material prevents the cracking that otherwise would result. The aseayer, who uses clay for luting, mixes with it for tlie same reason at least half its weight of sand. In the manufacture of fir^rick the required mixture is ground in a dry-pan, a machine similar in construction to the Carlin mixing pan (see Fig. 170), but provided with a bottom made of perforated plates to discharge the mAterial when ground sufficiently fine. Scrapers carried in front of the roUere throw material in their path, and the mixture when ground is screened, and further mixed in a horizontal pug-mill, being there tempered by the addition of water to the desired consistence.
The molding of brick is done by hand or by machine. If by hand the mixture is brought to the consistence of mud, and made into balls sufficiently lai^ to fill a mold. (See Fig. 20.) The mold is first sanded to prevent the adhesive mud from sticking, and this is thrown into the mcJd with force, to fill it completely, the excess is cut off with a stick or p,o, 20.— Brick-mould,
wire, and the brick dumped on a -
pallet or board. The pallets are placed upon racks, and air-dried until so stiff as to indent but slightly under pressure of the finger. They are then put through a re-pressing machine (Fig. 21), where they are given their exact form. When re-pressed, they are again placed on pallets and run into a dtyer which is divided into chambers and heated by steam, wastf
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heat or radiated heat, so that the last of the moisture is removed. The bricks, now so coherent that they can be handled with little damage, are piled in open order in the kiln, already described (see Fig. 19), and are burned at a temperature between 1230 and 1390° C, requiring one to three weeks for tlds.
In machine molding, called the " stiff-mud process," the clay is tempered with less water, and is much stifTer when molded than in hand-molding. The general form of the stiff-mud machine, known as the auger machine, is that of a horizontal cylinder, closed at one end, and tapering to a rectangular outlet, the size of the cross-section of the brick at the other.
Fia, 21. — RepreBsing Machine.
Within the (tylinder is a shaft carrying blades similar to those in a pug-mill, but at the end nearest the die, or outlet, the blades are replaced by a tapering screw. The tempered clay is fed into the cylinder at the end farthest from the die. It is mixed, and moved forward by the blades until seized by the screw which pushes it through the die. The bar of clay issuing from the machine is received upon a cutting table and cut into bricks by means of a wire frame. The further treatment of these bricks, with the drying, re-pressing, and burning, is like that of hand-molded bricks.
Another method of machine molding is called the dry-press process. In this method the mixture of clay and " grog " or coarsely ground brick is intimately mixed in a wet-pan with 10 per cent of water, molded in a dry
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Basin Refractories 37
preseing machine, and is then sent direct to the kilns for burning into brick. The expense of drying ia thus saved, but the brick is not of so good a quality as when otherwise made.
To resist abrasion, firebricks must be hard; to resist corrosion or Blag-ging, dense; and to resist high temperature and sudden changes of tempenture, porous and coarse in texture. We accordingly use the hard bricks for door-openings, dense ones for reverberatory furnace walls, and the porous and coarse ones for the roofs. The refractoriness of a firebrick depends on the quantity of the fluxing bases (especially alkalis) and silica conttuned, and on the coarseness of the grain. The grain depends again upon the degree to which the " grog " is ground.
Booe Ash. — This is made by burning bones, in a kiln wi^ an excess of air, and grinding the white residue to 20-mesh eize. Organic matter is thus removed and an impure calcium phosphate obtained. Iliough a neutral material, this resists the action of litharge, and it is accordin^y used, not cmly in assayii^, but in making the " teste " or movable hearttis of the English cupelling furnace shown in Fig. 278.
Basic Refractokibs
Dolomite. — The alkaline-earths, lime and magnesia, are strong bases and are resistant to basic slags, as shown later, but are readily fluxed by the silica of sihcious slags. Quick-lime is infusible, but is easily affected by the moisture of the air, and insufficiently coherent to be used for making basic brick. Dolomite is magnesian hmestone, and is a cheap refractory material. It is prepared for use by burning, much as is limestone. The proportion of lime to magnesia varies in dolomite, but the more magnesia the better for use as a refractory. The composition of a typical sample is as follows:
C»0 31.82
MgO 20, 19
SiO. 1 . 70
FeO 1.22
Co, 45.36
Dead-burned dolomite, specially prepared, has of late been substituted in part for magnesite as being cheaper.
Hagnesite. — This is the most valuable of the basic materials. When magnesium carbonate is calcined at a high temperature and dead-burned to 0.5 per cent carbon dioxide, the residue is practically infusible. It is used in grain form for furnace linings, or is manufactured into magnesite brick for the same purpose. Magnesite is usually colored dark-lm>wn l^
the presence of about 4 per cent iron oxide. It ie the preeenoe of the iron that enables it to bond or set well in furnace bottoms. Its main use is for basic open-hearth furnaces where the slag contains as little as 15 per cent silica. It is used also as a lining for forehearths (where it is in contact with low-^rade corrosive matte), also in lead, copper, or other heating or melting furnaces as well as for electric furnaces. The nature of the mineral is shown by the following analysis:
CaO 1,68
MgO 42.43
8iO, 0,92
Fe A and A1,0. 4-»
CO, Mid HtO 50.41
Carbon Blick. — Gas carbon, such as is used for arc lights, is made iiMo brick with a limited amount of gas-tar and burned in a kiln. "This brick has been found to be particularly resistant and refractory in a reducing atmosphere, as at the bosh of an iron-furnace.
Ottier Refractory Iflaterials. — A mixture of portland cement 2 parts, clay 1 part, and " chamotte " or coarsely ground firebrick 7 parts, moistened and molded into bricks or blocks, or used for patching furnaces, sets quickly and withstands a white heat without disintegrating. It iq early made and especially useful for rapid repairs. Only as much is mixed as is to be used at once.
While common red bricks are not refractory, the least fusible can be used in that part of the roof of a reverberatory furnace where the tem-> perarture is not high or only at a red heat. Such bricks are used for backing firebrick structures. As a general rule each kind of brick should be laid in a material similar to that of which it is composed. We should expect slagging to take place, for example, at joints made of loam-mortar in firebrick. Such loam, while cheap, is inferior to fireclay. An analysiB of good loam gives:
Percent.
SiO, 80 ,99
Total impurities 4.91
Ignition lose .- ,, 4,43
Here we note that the fluxing bases rise to nearly 5 per cent while alumina approaches 10 per cent, the ratio of the most fusible compound of alumina and silica. Where the fluxing bases rise above 5 per cent, there is risk of complete melting at high temperatures.
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Chapter Iv The Preparation Of Ores
We discuss this under the general heads of Sam{^u%, Cnishiiig, and Grinding, Screening and Claaaifying and Roasting.
We then take up the nature and operation of metallurgical furnaces and the principles of thermo-chemistry as preliminary to the whole question of roasting.
Principlbs Of Sampling
Sampling consists in obtaining from a large quantity of ore a smaD portion of a few ounces for assay. This must correctly represent the entire quantity of the ore, whether it be a few hundred pounds or thousands of tons, a wagon-load or a ship-load. Often we have a lot of ore, in which rich pieces mingle with poorer ones, or even with waste. In Sfunpling we must take this variation into account and represent each part, not only according to its value, but also to its quantity. Often ore is bought or sold upon the results of sampling. Thousands of dollars are involved and cash is paid for ore before the purchaser has treated it. In other cases, ores taken by the reduction works are treated separately, the owner receiving whatever is obtained, a charge being made to cover the cost of treatment and the profit to the reduction works. In this latter ease sampUng could be omitted. Similarly at a mill and mine, operated in one interest, the sampling may be omitted when considered an unnecessary expense. Efficiency of the work is their determined by the assay of the tailing.
If a reduction works is producing lead, copper, or zinc, in a form ready for market, the metals do not necessarily require to be sampled. Whenever the precious nsetals are also present in such quantity as to pay to separate them, however, the mettd is sampled to learn the values contained before selling to the refining works that is to effect the separation. In blast-furnace treatment, pre and all other constituents of the charge are sampled, assayed, and analyzed. From the data thus obtained, the charge can be correctly calculated and proportioned. . Not only is it necessary to ascertain the value of ores and of metals that result from metallut^cal operations, but as well -the value of the portions rejected. The efficiency of the work of the metallurgist depends
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Upon thorough extraction from the parts thrown away. To be assured of this, samples of slag or tailing are taken at frequent intervals. In finding the value of a lot of ore, we first weigh the ore, and base the assay value upon the dry weight. To do this we must determine the percental of moisture contained, as shown by a " moisture sample." We then sample the ore regularly, and finally assay the regular sample. Thus, suppose we have a lot of ore weighing 10,800 lb., containing 7 per cent moisture and by assay 54 per cent lead worth 3 cents per pound. Since the assay is made on the dry weight, we have, after deducting moisture, 10,044 lb. ore containii^ 5424 lb. lead worth, at 3 cents per pound, $162.72.
RBCEIVINO, SAUPLraO, CRUSHIHO, BEDDINO, AIlO STORIHO ORBS
The large smeltii^ works in the Rocky Mountain r^oo of the Western United States and Mexico buy their ores outright from mine-owners tor treatment. Such works are called custom works. A plant treating principally ore from its own mines is called a mine works. In custom works all the kinds of ores, already enumerated, are sampled and bought upon a schedule -of charges, generally established in advance between the works and the mite-owner.
Receiving and Weighing. — At reduction works that purchase ores (custom works), the ore arrives either loose or in sacks. Whether received by wagon or by car, the vehicle and ore are weighed together on platform scales, thus finding the " gross weight." When the vehicle is emptied, the weight, called the " tare," is similarly taken. The difference is the " net weight," or the " wet weight," and this is recorded. When ore arrives in sacks, the weight of the sacks also is deducted. Often sacked ore may be removed to scales to be we^hed, and only the wei^t of the sacks deducted, the difference being the net or wet weight. Sacks, if of sufficient value, are dried and returned to the owner. Railroads often return them without extra charge. Sometimes the sacked ore, if pulverulent, rich, or frozen, may be chained, sack and all, into the blast-furnace, the sack serving to retiun the fine contents until smelted, thus preventing the loss of flue dust.
The Moisture Sample. — In theory the moisture sample should be taken at the instant of weighing, since the ore may dry and become lighter. The sample is taken while the car is being unloaded or immediately after* ward. To represent by the sample the ore as contained in the car, holes are dug at average points (setting aside the dry top layer) and small poi^ tions are taken of ore that appears to be of average moisture. These portions are put in a covered can, and SO oz. <rf the mixture are weighed on a moisture-scale. After cautiously drying on a hot-plate, or preferably over-night on steam-coils, the 50-oz. portion is again weighed, and the
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' Hand 8Ampung 41
percentage of moisture deteimined by the loss in weight. Hie shipper often sends a representative to watch the samphng of his ore. Such a man should pay attention to this detail, otherwise too high a percentage may be deducted for moisture.
Sampling methods may be divided into two classes: hand-sampling and machine-or automatic-santplii^. Any method of sampling includes the starting and finishing operations.
Hand Sampling. — This includes the methods called " grab stunpling " and " trench aampUng," which are imperfect, and the r^ular methods known as, " coning and quartering," " fractional selection " and sampling with the " spUt sfaovd."
For determining the contents of fluxes and fuels and certain furnace products, the " grab sample " may serve. It consistB in taking at uniform distances over the pile or lot, similar amounts broken from the limits and taken from the fine. These portions are mixed and sampled by coning and quartering, by fractional selection, or by using the split shovel.
Coning and Quartering. — The ore is crushed and put in a circle or ring
about 8 ft. diameter on the samphng floor. The workman circles within
this ring, hovering all the ore to the apex of a cone at the center. This
completed with the shovel working from the apex radially, the ore is
drawn into a flat disk. This is marked by diametral lines into four equal
portions, of which two opposite ones
are left as I and III, Fig. 22, on the
floor and II and IV removed. The
reserved sectore are again shoveled
into a ring, then made into a cone, „ „ „ „ . „ ..
, ,, ., , ,, -_, Fia. 22. — Quartermg an Oro m Samplme.
now half the size of the first one.
This is again flattened into a disk, quartered and the two opposite quarters
reserved. The process goes on in this way until the sample has become of
small bulk, say of 2 lb. weight, when it should be again ground to pass
through an 80-mesh screen. It is thoroughly mixed by " rolling " on a
sheet of thin rubber cloth, and ttie nuxed product distributed into one or
several 4-oz. bottles or into manila sample-sacks which are marked with
the name and particulars of the sampled lot.
Fractional selectioo differs from the quartering method in that every second or fourth shovelful is reserved and coned as above described, for the purpose of mixing. From this cone each second or fourth shovelful is again reserved and coned, and this continues until it is necessary to recrusfa. After this, reduction in bulk again proceeds using a smaller shovel, according in size with that of the sample.
In sampling by ttte split-shovel, a good tool is shown in Fig. 23, the Brunton qiMirtering shovel, the central compartment holding the sample. From fljf ^JjfKKly'^nixed pile shovelfuls are taken, and by backward movements
The Preparation Of Ores
Fia. 23. — Bninton'e Quartering Shovel.
three-fourths of it slidee from the shovel blade into a heap, the remaining fourth in the central compartment being thrown into a separate pile. This pile is attacked in the same way, so that successively the amount to be sampled rapidly decreases.
The ore collected upon the sampling room Soor is there cut down by cutting and coning or by means of a Jones sampler, Fig. 24, until it we^hs 10 lb.
Fig. 24 is a view of a riffled sampler. The ore, evenly spread in the scoop, is so distributed in Fia. 24.^one9 Sampler. ^^^ "^'^ *^* one-half goes to
the rifi^t-hatid part and half to the left^and one. Either half is again cut down in the same manner, and so imtil but a small bulk remains, truly representative of the original.
Machine or Automatic Sampling. — It will be seen that the methods of sampUng by hand as just described, especially for large lots, involve much labor, and it has been
sought to overcome this by the use \
of machinery, A sample from a stream of ore, coming from a crushing machine, and called a " running sample," is taken automatically, this stream being deflected to one chute four-hfths of the time and to
Via. 25. — A Sample-grindiiiK Hill,
another, as a sample, one-fifth of the time, as shown in the Vezin e
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Sampling Mill 43.
j^, Fig. 27. It consists of a tube carried by a vertical abaft making 30 R.PJkI. '. Attached to the side of the tube and opening into it is a scoop. As the shaft revolves " counter-clockwise " the scoop (occupying one-fifth of the circumference) cute through the stream of ore from the inchned feed-chute for one-fifth of the time. The ore thus intercepted falls through the tube and becomes the sample, while the four-fifths, the rejected portioif , falling into the main hopper, is dehvered by the Fio. 26.— Braim Disk Grinder, chute to a bin.
SmnpHng MilL — Fig. 28 is a sectional elevation of a sampUi^ miU. An ore dump-car at the r^ht discharges its load into a sloping-bottom ore bin, whence it is drawn off t^ a sliding bin-gate and fed to a 15 by 9 in. Blake crusher (see Fig. 34). The discharge from the crusher falls into the boot of a vertical elevatqr, which raises it to the top of the building uid feeds it to the chute of a Snyder sampler or a Venn Sampler, Fig. 27, where 20 per cent of the stream is cut out to go to the 7 by 9 in. Blake cnisber just below. The . crushed product of the crusher is again spouted to a second sampler, where a{;ain 20 per cent of the stream is saved, and thence to rolls (see Fig. 43). The roll-product passes on to a third sampler, where again one-fifth or 20 per cent of the flow is caught to be more finely
Fio. 27. — Veiin Sampler. n l u i j . . n .v
roll, whose whole product falls upon the
floor of the sampling room. The rejected ore from upper sampler, amounting to 80 per cent of the whole, is shot back into an ore dump-car. The rejected portions from No. 2 and No. 3 sampler pass out at tiie side of the building to fall in a pile upon the ground.
The product is mixed on a mixing cloth, cut down by a smaller Jones riffle to 2 lb., dried in a steam-oven, and finely ground on a bucking-plate or in a disk-grinder, Fig. 26, to SO-mesh or finer.
The Biauu Disk-grinder ia here shown as opened for cleamng. Fine grinding ia done between the two fluted disks when the macliine is closed
Fig. 28.— a Sampling Mill.
for action. The Iiinged disk is pressed against the revolving one, the flutes conducting the ore downward between the grinding surfaces to the drawer
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Machine Sampuno 45
below. The ground material is passed through a screen, the oversize being returned to the grinder.
Finishing the Sample.— The groun<l product is now mixed by " rolling " on a rubber mixing cloth and distributed into manila paper sample-sacks that hold 3 or 4 oz. each.
The finaJ operations, from the taking from the sample safe, are called " finishing the sample."
At a custom works all the ore is stored in a sample safe or can and there is held until the ore-lot haa become the property of the works by purchase.
Sampling of Ores Containing Metallic Substances.^This is an operation requiring a clear knowledge of the principles of samplii^. We come upon these " metallics " sometimes in the operation of sampling. They muflt be separated, cut smaller, and quartered down separately by a hand-method, and reduced in size, at the same rate as the fine ore. If a fine substance ia made by cutting up the metallics it can be united with fine ore. Often metallics are brittle, but with diligent work can be broken, cut and " quartered down " without serious difficulty.
Cost of Saiiq>liiig. — In 1910 the cost of moving the ore cars, unloading into bins, returning the cars to the sampling-mill, and unloading the fractional part, usually one-tenth, retained was taken at 10 cents per ton. The cost of hand-sampling the tenth part was taken at 75 cents per ton. Hence, for ludoadit^ and hand-sampHng a 100-ton lot, the total cost was 17,5 cents per ton. At the Metallic Extraction Works, Cyanide, Colorado, ore was then unloaded from the car to a feed-chute crushed to J-in. size, automatically sampled and delivered to storage bins, for 11 cents per ton. A char^ of $1 to $2 per ton has been made for sampling, storing, assaying, and selling ore at custom works or sampling-mills, where the company haa acted as selhng agent and obtained the best possible price for the shipper. The price for sampling concentrate was 50 cents per ton less.
Sampling Concentrate, Tailing, and Ore-pu^. — C-concentrate is sampled easily, for it can he thoroughly mixed and sampled by hand. TaiUng carries .40 to 50 per cent moisture. It has but little value, and needs no close attention. Ore-pulp flowing in a launder is often automatically sampled.
Fig. 29 shows the machine used for this. The pulp flow, entering by the chute on the left, ia carried away by a launder set to receive it. As the wheel revolve^s, the attached scoops divert part of the flow into a compartment of the receiving box. When not so sampled, a bucketful, taken each hour from the stream, and all these samples united in one portion after decanting the wat«r, may be used to determine the approximate daily average value. When loaded into car^, the sample is sometimes taken by boring to the bottom of the body of ore, using a ship auger, or again by a pipe driven downward. This is a kind <^ grab-sampling, but
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serves in mills where such an approidmatioii is considered sufficiently
accurate.
Sampling Iron Ons. — These, being uniform in constitution, are more simply sampled. The railroad cars are sampled by taking a grab-sample at six or eight places uniformly over the load. From this tm analysis determines the character of the ore, and where it is to be stored for shipment by ore-boat. Lat«r the cargo is sampled at the receiving port. This is done by taking grab-samples upon the exposed surfaces of the ore in the hatches while it is being imloaded the ore being immediately put in closely covered cans so that its moisture shall be conserved.
Mill Samples. — The value of the ore goii^ to the mill (the mill heads) is estimated as the value of the product plus the tailings loss. As a check a sample is taken of
F,a. 29.-Mi„„ aunpter. '>" ("^-""^ ^h^, this bei«g
dried, solution and all and assayed. Knowing the specific gravity of the discharge we can compute ita contained solution, whose value is subtracted from that of the solution taken as a drip from the storage solution tank. This checks within 4 per cent of the product-plus-tailings result.
Dip-sanqtles are taken hourly of the various agitator and thickener pulps and solutions. To prevent further dissolution by the cyanide, 10 c.c. of a 10 per cent solution of sodium sulphide is added to each dip-solution once per shift. They are filtered, washed and dried for assay, and results promptly reported.
Principles <rf Sampl)i)g.^In the progressive crushing above described, it will be observed that the ore is made finer as the sample becomes leas. This is to make sure of a constant ratio between the size of a single rich piece and the whole sample, that such rich piece shall not produce an appreciable effect on the assay value, whether it be present or absent. The richer, and at the same time the more " spotty " or varied the ore, the finer it should be crushed before cutting-down or quartering. The table below shows how this is arranged in practice:
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Machine Sampung
Size Of Largest Pieces
v.... .H fl...„ o.»c„
T» Ton.
Wsicht or On.
HIsbat 3000.
Cocoanut
Orange
Walnut
Pea
SO-mah
Fiet
Egg
Cheatnut
Wheat
lOO-meoh
We may conclude that for accurate sampting the requirements are :
(1) The takiBg uniformly frequent portions to ensure an average of the stream of ore aa it is undergoing progreeaive sampling.
(2) Thorough mixing of the ore to ensure uniform richness.
The Martin Sampling Machine. — This machine excavates the slime or mud of flotation concentrates containing 20 per cent moisture from railroad cars transferring it to bins and taking out a sample in so doii^.
The bin, long and narrow, has a slit 4 ft. wide in the bottom covered by tnmsverse plank. The furnace charge-car runs beneath the bin, so that the material is drawn off to the car by taking up the planks progressively. Centrally above the bins is the sampling machine which travels freely delivering the slime where desired.
On the nearby parallel track is stationed the train of cars loaded with the slime. The machine is furnished with a grab-bucket, and moves along the track where needed for excavating. The load is delivered upon a slowly moving 8-ft.-wide endless sheet-metal belt discharging thence upon a reel the width of the belt. The reel is made up of several disks 18 in. diameter with wires stretched from end to end over them at 6 in. intervab. The sheet of slime is chopped up by the disks and wires of the reel and falls in lumps upon a hollow cylinder of 15-in. diameter having a sht in it of one-tenth it^ circumference. As the cylinder revolves the material falls upon it, and when the slit comes under them it falls out of the sUt again upon a conveying belt covered with a dry calcareous sand from a nearby deposit. This belt takes this supply from a bin, the head pulley of it being within the travelir^ sample mill and dehvering its load of sand and slime where another divider belt takes out a tenth, so that a 1 per cent sample is obtained.
It will unload 80 tons per hour, where, when unloading by hand, the cost was 20 centa per ton.
Pulp Sampler. — As shown in Fig. 30, this consists of a disk with tai attached pipe. The disk is counter-we^jhted, so that when past the axis,
The Preparation Of Ores
it suddenly falls, the pipe sweeping through the tailings stream. One may note a slot in the pipe where the pulp from the tailings stream enters. As
the pipe is lifted with the disk again into its vertical position the pulp flows out of the lower end of the pipe into the sample launder.
Sampuhg Ubtals
Metals may be sampled either in the solid or molten state.
Gold or Silver Bars or Ingots.— These are sampled for assay either by granulating a small portion of them or by taking chip-samples from them. In the first case, while the metal b in a molten condition, a small ladleful, weighing an ounce or less, is taken from the crucible immediately after stirring it. This is poured into a bucketful of water, thereby granulating the metal and forming particles of a variety of sizes, convenient for weighing and assay. Chip-samples arc taken at points diagonally opposite on the edges of the bar, and a cold-chisel, cutting out a small wedge-shaped piece, is used for this purpose. The pieces are annealed and rolled into a ribbon for assay. The average assay-value of the two pieces thus obtained is taken as the true value.
Base-bullion. — This b lead that comes from silver-lead blast-furnaces, and it contains commonly 100 to 400 oz. of silver per ton. When poured into molds to be cast in bars, the silver segregates, and the exterior of the bar, that cools first, is richer in silver by several ounces than the central part. This is illustrated in the cross-section of a bar (Fig. 31), in which
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SAMPUNG METAia 49
the center of the bar assays 10 ounces less than the exterior. Base-buUion
is sometimes sampled by taking two " chips " or
punchings, one from the top and one from the bottom
punch, is 8 in. long and removes a cylindrical piece of
Prom a carload lot of 400 bars, 800 of these chips
would be obtained. These are melted and the fused *t«,i« au ». ^i.., p« fa»
metal stirred in a plumbago crucible and cast into yw. 31.— Distribution
bar. This is a sample of a 400-bar lot of base-bullion, of Silver in e, Bar ol
equivalent to 20 tons. A better way of sampling, Base-bullion.
however, is to remelt the metal in a large kettle (see
Fig. 276), and to skim and recast into bars for shipment. While casting the metal, a sample is taken from the molten bath and poured into a bullet^mold of such a size that each F,a.32.-B,«-b„ll,onSamplmB-pmch. ^^„^^ ^^.^^^ approximately a half
assay-ton. This is trimmed to the exact weight for the assay.
Copper Ingots or Anodes. — The s€^;regation of gold and silver in copper ingots is even more marked than in bars of base-bullion. This is shown in Fig. 33, which represents the distribution of gold and silver in an ingot of blister-copper 5 in. deep. In this case, however, the interior is higher both in silver and gold. The usual way to sample such bars is to drill into them and retain the borings for a sample. Manifestly a sample hke this is uncertain, and depends upon the selection
of the place on the ingot for takmg p,^ 33._Section of Bar of Ingot it. To obviate this difficulty, in sam- Copper,
pling a lot, say of 100 bars, it is customary to drill into each succeeding bar at a different spot to obtain an average by so doing. It is preferable, however, to take the sample at the time the copper is melted and well mixed in the furnace by poling. As in the case of base-bullion, samples of copper are taken while dipping or casting, one at the beginning, one when the charge is half removed, and one toward the end. The average of the three samples is regarded a correct representation.
Pig Iron.^ — This is sfimpled and graded by inspecting its fracture and l^ chemical analysis. When an analysis is to be made, the sample is taken from the drillings of a small bar, molded while the metal is flowing from the furnace. The percentage of silicon determines the grade, the whole oi the sample pieces being dissolved for assay.
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Chapter V
Crushing And Grinding, Screening And Classifying
Size of Run-of-mine Ore. — This, as it comes from underground, or from open-cut, will vary from pieces several tone in weight to dust which, when wet, is called sHme. The larger pieces may be broken to sizes suited to loading by band sledging or by block-holing, that is, by drilling and shooting them. Since this is more expensive than breakir^ by rock-breakers, the plan is to have these machines large enoi^h to take such pieces, especnally if they are hard to break.
PRINCIPLES OF CRUSHmO
Power Needed in Crushing. — The work done in crushing varies Inversely with the size to which the ore particles are crushed, or directly according to the increase of surface.
Thus a ton of quartz ore, composed of 1-in. cubes, if crushed to 0.5-in. cube would need 0,257 horse-power to so reduce it. The diameter of these smaller cubes will then be one-half, while their surface will be double that of the original ones. The surface of a 1-in. cube is 6 sq. in., and that of the resultant eight 0.5-in. cubes will be 8X6X0.5*= 12 sq. in. Were the 1-in. cubes crushed to 0.25 in., or one-fourth, the surface of them would be 24 sq. in. and the power needed would be 0.514 H.F. This rule is fairly constant with the coarse sizes, but with the finer ones the increase of surface is more rapid than the work needed to produce them.
Action of Machines in Crushing. — Crushing may be effected iq either of two ways: In the first, the breaking is due to the impact of two approaching surfaces, as in rock-breakers, stamps, or rolls. In the second the surfaces move over one another, the ore being interposed and abraded or sheared, as, for example, in grinding pans, coffee-mill grinders, and the BrauQ sample-grinder. Often both actions are taking place, as in ball-mills and tube-mills.
Stage Grinding. — The operations of reducing an ore to a fineness, such that its mineral particles can be acted on by a solvent, separated by concentration or roasted, is done by crushing in stages. The first st^^, or coarse-crushing, is done by rock-breakers reducing the ore, so that its larger piece will be no more than 1.0 to 2.0 in diameter. The second, or SO
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intermediate cruBhing, aims to reduce the coarse-crushed ore to 30-or 40-mesh uze, this being done by stamps, rolls, or ball-mills. The Uiird, or fine grindii^, takes this product and grinds it, so that much of it will pass a 200-mesh screen.
A point to be observed in preparing <»« for leaching is to avoid malting slime in crushing. Any considerable portion of finely ground or slimed material hinders percolation greatly. If grinding is performed by rolls, a more granular product, containing less dime, is produced than by crushing to the same screen-size with stamps. If it is desired to obtun the maximum quantity of sand and the minimum of slime, then gradual reduction, or graded crushing, should be adopted. This consists in first screening out the ore, already sufficiently fine.
Screen Sizes. — We specify the size of a piece of ore by saying that it win just pass throw^ a 2-in. or 4-in. ring, for instance, or through a screen having round or square openings of that size. For smaller pieces or particles, the size is designated as being able to pass through a wire-mesh screen, the distance, center to center of the wire, being meant. The opening is less than this distance by the thickness of the wire. Thus a 20-in. mesh screen would be one having 20 wires to the inch, or 0.05 in. center to center, and also 20 openings, whose size would depend on the thickness of the wire. Thus, the diameter of tiie wire might vary from 0.025 to 0.009 in., and the resultant opening from 0.025 to 0.041 in. Where the thickness of the wire is equal to one-half the mesh, then the percentage of opening is 25 per cent, and the largest particle that can pass the screen is iSr, or 0.025 in. This proportion is sanctioned when the actual wire size has not been specified.
COARSE OR PRIMARY CRUSmnG
Two kinds of rock-breakers are in general use for coarse crushing — the Blake and the gyratory. Blake crushers, with a feed-aperture of 5 ft. by 6 ft., have been made so that pieces of rock of nearly that size, and weighing four or five tons, can be broken. Such large machines are expensive, and so hand-breaking may be done when the rock favors it. Blake crushers have been made to take a boulder as large as 5 ft. while recently, gyratory crushers have been built having a receiving opening of 5 X 15 ft., weig^ng 237 tons, and having a capacity of 2500 tons per hour. For the same width of aperture a gyratory crusher costs nearly two and a half times the Blake. It is convenient to remember that either kind requires about one horse-power to cnifh one ton per hour.
The Blake Crusher. — Fig. 34 is a view of a Blake crusher or breiJter. ^. 35 is a section of oae having a receiving openii^ 20X20 in. This cruriies ore as it comes from the mine, containing pieces as large as 12 in. diameter, at the rate of 25 tons per hour. There are, however, <s\ay9y
52 Crushing, Grinding, Screening And Classifying
and talcoee wet ores containing 25 to 30 per cent moisture that stick to the rock-breaker, and are impossible to crush in the wet state. Such ore may be first dried in a cyUndrical dryer. The Bhike rock-crusher is shown in perspective in Fig. 34 and in longitudinal section in Fig. 35. It con-wtB of a heavy cast-iron frame, marked JV, within which is placed the fixed jaw F and the swinging jaw B and between them the ore is crushed. A shaft T, eccentric where it passes through the pitman K, causes this to rise and fall, producing a corresponding movement of the adjacent ends of the toggles /, /. As these rise, the effect is to push the jaw forward to
Fio. 34.— The Blake Ore-cniBher.
produce the crushuig movement. As the pitman and the tofO^lcs descend the jaw recedes, and is pulled baek by the spring rod P and the spring R. Flywheels A help to steady the movement. The machine is driven by the pulley at 250 revolutions per minute. The movement of the lower end of the jaw is J to | in. For the breaker above specified, the discharge-opening would be 20 by 1 J in. to crush to 1 i-in. size. The receiving opening would be 20 by 12 in., and would take pieces as lai^ as 12-in. diameter.
The Gyratory Crusher. — Fig. 36 ia a perspective view, and Fig. 37 a sectional elevation of this type of crusher. Referring to Fig. 37 (the bottom plate dropped) is a main frame or body 2 and 3, witii a three-
Coarse Crushing
FlQ. 38. — Gyratory Crusher.
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M Crushing, Grinding, Screening And Classifying
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OOAfiSE GRINDING 56
legged apider to carry the top of the spindle or vertical shaft. This is surmounted by an ore-hopper 7, the ore taUiog between the legs of the spider to be crushed between the hners, 19 of the body 3, and the cone-shaped head of the spindle 25. The lower end of the spindle is moved in a circle without revolving, by an eccentric sleeve 8, made in one with the bevel gear 9. Thus the opening between the head and the liner is alternately opened and closed, crushing the ore, the product discharging over the chilled weariog-plates 22. The amount of the jaw-opening can be varied by raising the spindle using the lighter-screw 29. At 12 is the belt-driving pulley.
DfTBRHEDIATE OR FIRE CRUSHING OR COARSE GRDfDIItG
This is done largely by stamps, by pans, by ball-mills, by rolls, by Chilian mills and by the Symons disk-crusher. Ball-mills and pans are
good intermediate grinders up to 100-mesh, and to this point they are economical. There is a tendency at present to supplant stamps, heretofore so largely used, by ball-mills, except in cases where inside and outside amfdgamation is to be used.
The StanqMnilL — This is given under the bead of amalgamation as uaed in gold-mill practice. It will take ore of 1 to 1^ in. in diameter and will reduce it to pass a battery-screen aperture, economically of not less than 0.15 to 0.4 in. The stamp-mill has reached Hb greatest development on the Rand in South Africa. It is held that while heavier stamps have been used, those of 1600 lb. falling weight are the practical
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Crushing, Grinding, Screening And Classifying
limit. The tendency in American practice is to replace the stamps by ball-mills in new construction.
The Ball-mill. — Pig. 38 is a view of a ball-mill, showing the spiral scoop that picks up the feed and delivers it into the mill and the
Fio. 40.— Ball-n
n Ooscd Circuit.
flaring discharge for the exit of the groimd pulp. The discharge is screened to remove the finer sand, while the coarser, dischai^ing at the extreme end, is returned for rcgrinding. The mill has a self-locking lining needing no bolts through the (iylindrical shell. The mill is filled about half full of flint-pebbles, such as are employed for tube-mills, generally imported from Scandinavian countries or from France. In place of pebbles, iron or steel balls arc increasingly in use, these often toughened by the addition of manganese or chromium. The dischai^ of the mill may be protected by a grid or perforated plate, see Fig. 39, as in the Marcy mill, to hold the balls or pebbles back while permitting the escape of the ground material.
Proportioiis and Efficiency. — Fig. 41 shows the paths of travel of particles in a ball-mill 8 ft. diameter by 6 ft. long. It is assumed that the - steel balls are from 3 in. to 2 in. diameter, and that when they are much smaller than this they are
Fio. 41 . — Pal li iit Travel of Ore-particles.
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Coarse Grinding 57
removed, as beii^ a hindrance rather than an aid in comminution, being replaced by S-in. balls. At the correct speed 22 K.P.M. there will be needed 180 H.F.
On a medium-hard porphyry ore, such a mill should grind in twenty-four hours from 2j in. size to minus 4&'me3h about 45 tons, or of a hard quartz ore half, and of soft porphyry twice this. We distinguish between the circulating load, .which is that which makes the cycle through the mill and classifier, and the input load, being the original ore plus the circulating load. The water used will vary from 30 per cent to 50 per cent of the ore for the most efficient grinding. The rniW is lined throughout with heavy manganese-steel plates.
The Hardinge Conical ACilL — This ball-mill, as shown in Fig. 42, has conical ends with a cylindrical center of 36 in. for a mill 8 ft. diameter
Fiu. 42. — Harilinge Conicul Mill.
by 22 ft. long. It feeds and delivers as in the Marcy mill. Due to it« shape, the larger balls keep to the larger diameter, so that only the finely ground product and the worn particles from the balls discharge. Elach day a few new balls are fed into the feed-scoop.
For best performance in a mill of this size the balls would weigh 28,000 lb. and would occupy 0.3 to 0.4 of the volume of the mill, or as much as is shown in the figure. It should yield VJ tons per hour on average ore, covering from J in. to 20O-me8h size.
Fig. 40 is an elevation of a Marcy ball bill, 4^ ft. diameter, inclosed circuit with a Dorr classifier. At the right end is a feeder, which scoops up the coarsely crushed ore and the water from the box which receives this feed from a launder. Entering the mill, the ore is crushed by the touKh manganese steel balls that are lifted up and fall upon it, due to the revolution of the mill. The mill is driven by herring-bone gearing and electric
58 Crushing, Grinding, Screening And Classifying
motor at 230 R.P.M. The disch&i^ goes to the lower end of the Dorr classifier. Here the ui^ound particles sink and , are gradually raked to the upper end to return to the mill-feed box. At the lower end of the
Fia. 43. — Fifty-four Inch Crushing Rolla.
classifier the floating, finely ground product or slime overflows, ready for further treatment.
Crushing Rolls. — We show in Fig. 43 a perspective view of .rolls having grooved shells or tires; in Fig. 44 plan and elevations of rolls having smooth shells. In the plan and outside elevation in Fig. 45 the fixed roll is at the right, in Fig. 43 at the left, and the movable or spring-roll is carried in sliding boxes. The long, heavy tension rods have capstan nuts at both ends for i tightening the springs against the ' sliding boxes. In case a hard object get« into the rolls as they revolve, such as a hanuneivhead, the springp give back, permitting its passage, so that the rolls do not stall. In Fig. 45 the roll-shell or tire is pinched between the fixed roll head 7, and the movable one 8, due to a slight taper on the heads corresponding to a like one on the shell. Bolts draw those heads tightly together. For heavy work
Co Otitic
Pia. 44.— Angle of Nip.
Coakse Grinding 69
such rolls have been built up to 72 in. diameter by 24 in. face. They will nip even to 4-in. pieces, and with a " choke feed," that is, with a thick
Btream of ore fed on to the rolls while well set up, will run smoothly with a capacity of 2500 tons per day. Due to ita grooved shell, the smaller rc^, Fig. 43, will take as large a piece.
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60 Crushing, Grinding, Screening And Classifying
Present practice in roll crushing involves a high peripheral speed, not only as increasing capacity, but also to insure smoother running, owing to the fact that the inertia of the rolls carries them safely by a sudden peak of load. Such speeds are from 300 to 500 ft. peripheral speed or, for 42-in. rolls, from 27 to '45 R.P.M. It is the custom to drive the rigid roll by a large pulley, using a smaller pulley on the spring-roll. The fixed or rigid-roll shaft has a deep groove turned at one end of it, fitted with a thrust-bearing. By means of two bolts this thrust-bearing may be moved axially, and with it the shaft. This results in giving another surface of
Fia. 46.— Chilian Mill.
contact between the rolls, preventing them from grooving. In some makes of rolls this end movement, called " floating " is slowly and automatically performed. Rolls are also made without springs; they are called " rigid " rolls. They have the atlvantage that they produce less fines and practically no oversize. They run with little jar or vibration. The size of feed for 42-in. rolls would be 1^ to 2 in. and the reduction of the size of the feed is preferably four to one, that is 2-in. pieces should be reduced to J in. size.
The Chilian Mill.— Fig. 4fi is a view of a 5-ft. mill with the shrouding broken away to show two of the three rollers. These have steel tires. They roll upon a die-ring. Ore is fed by launder to a central hopper set
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Coarse Grinding 61
above the roll axles, it flows downward into the pan below and outward so as to be ground between the rollers and the ring. It is splashed outward, and when ground fine enough, escapes through the peripheral Bcreens that form the sides of the pan and into a gutter by wUch it flows away. A 5-ft. mill, running at 40 R.P.M., will grind 25 to 35 tons of ore in twenty-four hours, requiring 10 H.P. to drive it, and usii^ 400 to 1000 gallons of water per hour.
Fia. 47. — SymoDs Disk-crusher.
Besides the fast-running type of Chilian mill above described there is the slower kind, called the Lane, running 12 to 15 R.P.M. and having rollers 7 ft. diameter by 22 in. face, and weighing 7 to 8 tons each. These travel on a die ring 7 ft. diameter and crush through a 30-mesh screen at the rate of 15 tona in twenty-four hour«.
The Symons Vertical Disk Crusher. — We show in Fig. 47 a pei> Bpective view; in Fig. 48 is a longitudinal sectioa of a 48-in. Symona disk grinder. Referring to the section Fig. 4$, there are two saucer-
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Crushinq, Orinding, Scbegninq And Classiftino
shaped disks carried at the end of a sleeve and with a slit opening revolving at 100 R.P.M. Within the sleeve ia a shaft whose tail-end receives a reciprocating motion, due to an eccentric with its pulley, at 250 R.P.M. Where the disks are the widest apart the falling ore from the feed-chute enters and is pinched and shattered as the slit closes, the fines being thrown out into the housing and falling to a discharge^pout beneath. The machine is not suited to grinding sticky ores, which tend to adhere to the disk. Friable ores work well in it. Another defect of this crusher is that " tramp iron " is liable to break, or stall the machine. This is overcome by the use of magnets at the conveyor belt.
The Degree of Comminution of the ore is one of the moat important factors, particularly in the treatment of silver ores. The piupose of grinding is to free the minerals from the incloBing gangue and to reduce the mineral particles to such a size that they are readily dissolved. Grinding should be carried on in such a manner as to waste as little power as possible in grinding the worthless gangue and still fulfill the above-named conditions to the fullest extent possible, since the less the mineral is protected by the gangue and the greater the surface exposed to the solution, the more rapid the rate of dissolution.
It then follows, for example, in the case of certain silver ores, that while fine grinding may not produce any greater ultimate extraction, yet in general it will materially reduce the time of treatment necessary. But this advantage is not always reaUzed without the disadvantage of greater consiunption of cyanide arisii^, since finer grinding not only causes a greater surface of the minerals containing the precious metals to be exposed to the solution, but also a greater surface of those minerals, if present, which may act as cyanicides. Careful correlation of the time of treatment with degree of comminution may serve to minimize this difficulty.
Selective grinding whereby the heavy mineral particles are ground finer than the lighter particles of gangue takes place automatically to a greater or less degree in closed circuits where hydrauUc or mechanical classifiers are empl(^ed so that actually in the majority of plants the heavy mineral particles are ground finer than the gangue. The additional cost of finer grindii^, together with the attendant disadvantages which may arise, must in each case be carefully weighed against the additional extraction I)ossible, or the decreased time of treatment necessary to obtain a given extraction.
In grinding a sticky talcose ore, the pebbles may become coated, the noise dies down and grinding ceases. If, however, the feed is cut o£f for a
Fig. 40.— DiekH of Symons Crusher.
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Fine Grinding 63
few nuDUtes the pebbles free themselves, the noise resumes, and grinding again continues.
Ban or Tube-mill Drive. — Fig. 50 is a plan of an excellent drive, because of the use of a flexible coupling. It does away with the trouble
Fio. 60.— Plan of Ball-mill Drive.
due to the strains on the transmission. As arranged it drives a 6-ft. by 4i-ft. ball miU at 24 R.P.M. ; a 7 by 12-ft. tube or pebble-mill at 22 R.P.M. Tube-mill Linings.— To withstand the wear of rolls or pebbles on the interior of a ball or tube-mill, special tough steel or iron plates are provided. In Fig. 51, A represents the Tonopah lining where longitudinal
FiQ. 61.— Tube-mil! Lineio.
ribe give spaces to be filled with concrete. B is the so-called M Oro liner, shown with its load of pebbles while revolving. The pebbles after a few revolutions wedge themselves into the grooves, affording a resistant surface to the action of the falling pebbles. In C we have the Komata liner made with heavy ribs. These, during the revolution oi the mill, raise the pebUes quite high before they can fall back.
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Crushing, Grinding, Screening And Classifying
FIIfE ORINDINO
Tube Mills. — They are the best of the all-sliming machines, and are commonly in use in modern plants, either for combined sand leaching and slime treatment in cyanide practice, or for all-slime treatment. F^. 52 gives a view of such a mill. They have been largely made 5 ft. diameter by 22 ft. long, but now mills 6 ft. diameter by 14 to 16 ft. long are preferred. The interior is lined throughout by thick steel plates which last from nine to twenty-four months. The method of the feed and dis-chai^ is the same as for the ball-mills.
The speed varies from 28 to 34 R.P.M. The cylinder is filled half full of hard flint pebbles. The falling and roUing of the pebbles on each other and upon the lining as the cylinder revolves grinds the material, in part by impact, in part by rubbing abrasion. The pulp, fed to the mill, varies
bio. 62. — Tube-mill.
in' size from J in. diameter to 40-mesh size, and preferably should contain 40 per cent water. In sliming to minus 200-mcsh, the machine is worked in closed circuit with a classifier, which allows only the finer pulp to escape, and returns the coarse niatorial to be reground.
Dry Crushing and Screening. — Ores are dry-crushed and screened in a closed circuit as a preliminary to roasting. It is done by means of rolls and revolving screens or trommels in series, as shown in the diagram or flow-sheet. Fig. 53,
The ore supply from a dryer such as the Whitc-Howell roaster. Fig. 73, goes to the roughing rolls a which reduce it from 0.75 to 0.25 in. The crushed ore is raised by the elevator to a trommel, or separating-screen, having screens of J-and j-in. aperture, respectively. The first two-thirds of the screen takes out all material less than J in., and the final size is all coarser than J in. We thus get three products, an oversize from the coarser screen which goes back to the roughing rolls to be rccrushed, a screened product or undersize, which goes to the medium rolls, b, set at \ in. open, there to be crushed and sent back to the separating screen, and finally
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Stage Grinding
65
an undersize through i-in. screen, fine enough to go to the finiwhing rolls. Until crushed so fine that it passes the finest mesh the ore is returned to the trommel. The fine product of the screen is raised by the elevator /' and the ore stream is equally divided between a" and a"', which are provided with 30-mesh wire cloth. The undersize from these trommels drops into the storage bin m, while the oversize is conveyed to the finishii^ rolls, after which it goes by elevator to the finishing screens. Thus, nothing enters the bin except SO-mesh, or finely crushed ore ready for further treatment.
This system of graded crushing is preferable because the final product
Fio. 53. — Plow-flheet for Dry-nruahing or RoreeDing.
contains a minimum of fine or slimed ore and being granular is more easily percolated or leached. As each piece or particle of ore is crushed by a single nip, \he fine is separated by the screen and protected from unnecessary breakii^ with consequent waste of power. The chief coats in thb ^yBtem of dry-crushing are those of labor, power, supplies, and repiurs. These vary with the tonnage. The cost of crushing in 1913 to SO^^nesh size, in preparation for roasting or leaching is 50 cents per ton, but to this must be added overhead or general expense.
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Crushing, Grinding, Screening And Classifying
Scseenihg
Screening. — A mixture of coarse and fine mineral, shaken through a baT'-Bcreen, a plate-screen, or a wire-cloth screen will divide into two products, an oversize or coarse product, which remains upon the screen, and an undersize or minus product which drops through its openings, and which is of the size of the opening down to the finest particles of the ore. The oversize above mentioned may be again crushed and fed upon the screen, and by sufficiently repeating the operation it all eventually will pass through as imdersize. Or again, the undersize of a screen may be brought upon a finer one and will then yield an oversize product, which will vary from the minus size of the first screen to the plus oversize of the second. Thus, by using a succession of screens, we obtain a series of products, grades from the coarsest to the finest, and a residual one of all finer than the finest screen.
Classifsnng. — Finely ground ore is called " pulp " — ^if dry, a " dry piilp," if wet, a " wet pulp." It consists of coarse and fine particles called respectively sand and slime. Such materials, when wet and suspended in a hquid, may be separated into sand and slime by means of hydraulic classifiers, the sand settUng and constituting the " underflow " or " spigot " discharge of such an apparatus, while the slime, remaking in suspension, passes away above at the " overflow." Classifiers may be of large capacity, and are cheaper and more durable than screens.
They may be divided into two kinds, viz., hydraulic classifiers and mechanical classifiers.
Bar Screen or Grizzly.—
This is made of bars 6 to 12
ft. long, spaced about li in.
Fig. M.— Griiriiea. apart, Fig. 54. Setatast^ep
slope, as in a, Fig. 134, it
receives the ore dumped upon it from the mine-car. The finer ore drops
through the bars to the bin beneath, the oversize joins it after passing
through the crusher.
Tlie Impact Screen. — For the dry screening this screen. Fig. 55, is
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Scbeenino
well suited to fine screening. The screen frame, highly inclined, receives a bouncing or bumping motion, thereby giving large capacity in a small
Flo. 65— The Impact Screen,
PiQ. 56— The Tension Screen.
Space with a minimum wear of cloth. The launder at the head has pointed cleats or guides which distribute the feed evenly over the screen.
08 Crushing, Grinding. Screening And Classifying
The Rev<^Tiog Screen or TrommeL — In Fig. 57 is a trommcj, mounted on a shaft inclined coinmonly 4° to 5° to the horiKHital, and revolving to 20 times per minute. The cyUndrical part b covered by a punched plate or by wire cloth. . It has at the left a short receiving cone where the ore enters a V-tihapcd shcct-^teel housing or casing receiving the watery undersize, which discharges by the spout A. A second spout having a dividing partition to prevent mixing, discharges at B. Any portion of the ore may Iw said to travel throujdi the trommel in a spiral path. The trommel is operated wet (sometimes dry), wash-water being fed on the out«de of the up-eoniing side by a spray pipe. A practical size for a trom-
Fio. 57. — Il<!volving-epreen or Trommel.
mel would he 3 ft. diainotcr by 6 to 8 ft. long with a capacity through i-in holes of 200 tone per day.
Classifying
Classifiers. — These may be dividetl into two type,4, the hydraulic and the mechanical classifiers. In the first we include the Caldecott and the Allen cones; in the stx^ond the Dorr and the Aldns classifiers. Of the cones, the Caldecott needs much attention, in order at all times to produce a uniform food to the tube-mills. It returns much fine material and ia a continuous source of trouble. The .\llen cone, being automatic in ts action, is not nearly so open to these objections. The Dorr and Akins classifiers require little attention, and will deliver a feed with 20 to 30 per cent moisture. The chief objection to these machines, when dealing with a heavy sulphide concentrate, is the delivery to the tube-mill of a considerable amount of fine material that docs not require finer grinding.
Classifying
The Caldecott Diaphnigin Cone-cUs^er. — This hydraulic classifier, Ji^g. 58, 6 ft. diameter by 9 ft. high, is popular in South Africa for supplying & feed to tube-mills. Its advantage over the ordinary cone type is that it has a diaphragm which sustains the sands accumulating above it in the ttaik nearly to the inlet spout as shown in the view of the Allen Cone, Fig. 59, while at the bottom is a cut-off valve, that can be adjusted to r^^late the discharge. The pulp feed of the launder passes throu^ a screen to remove clips and floating objects, and falls in the cylindrical inlet spout upon a deflector to gently settle out the sands. The slime overflows over the entire circumference of the cone into the peripheral launder.
The Allen Cone.— This classifier. Fig. 59, like the Caldecott, is of the same settling-basiu type, not using hydraulic water, the sand settling in the basin Fia. 68.— The Caldecott Diaphisginwhile the slime escapes by the extended cone,
peripheral launder. The settling solids,
forming a sand-bed, are removed as they accumulate by an automatic float-controlled spigot-valve 29. The feed drops downward through the inlet pipe or spout 1, impinges on its casting, which quiets the flow, and seeks exit through the truncated cone 3. When the sand-bed accumulates to near the mouth of the cone it partly obstructs the outward flow, causii^ the watery pulp feed to rise in the cone and carry with it the float 5. By means of the lever 13 the link 28 actuates the ball valve 29 and the settled aoUds, even as large as i-in. pieces, begin to flow out until the level of the solids in the basin drops, and with it the float 5, again permitting free flow. The superiority of this cone over other hydraulic classifiers is in the automatic control of the sand dischai^, since, when properly fed and adjusted it needs no attention.
The Akins Classifler.— In this machine, F^. 60, as in the Dorr classifier, which follows, the sand is elevated from an inchned settling tank or trough by mechanical means, in this case by a spiral or screw, which fits closely to the semicircular bottom of the trough, eccaping the sands upward to the dischai^ at the upper end of the trough, while the slime is agitated and overflows a diaphragm at the lower end to the discharge pipe R.
The Dorr Boiri Classifler. — This is a Dorr classifier to which has been added a bowl or tray, as shown in Fig. 61, which receives the
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70 Crushing, Grinding, Screening And Classifying
Fin. 59.— The Allen Cone.
Fio. 00.— The Akina ClaaBifier.l <_^
Classifying 71
feed at its center, the slimed pulp overflows the periphery into the overflow launder, while the sand is plowed to the center of the bowl and discharged through a comparatively small opening into the main inclined settling tank. Here it b raJced to the upper end of the tank by means of two sets of reciprocating scrapere. The sand settling to the bottom is gradually advanced up it by the forward motion of the scrapers. The mechanism then lifts these scrapers which drop to the bottom, then make
Fio. 61.— The Dorr Bowl-claasifier.
the next forward stroke. The coarser material {the sand), emerging from the solution ia discharged at the upper end still containing 25 per cent of moisture. Wash water or solution is admitted into the main classifier tank and flows underneath the bowl and up through its central opening, counter-current to the sand. Usually about 1 ton of back-flow wash per ton of sand is sufficient to remove all slime from the sand. The agitation is just enough to prevent settling of the alive which now overflows at the periphery of the bowl.
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Chapter Vi
Uetaixurgical Furnaces
Classification. — These may be divided into two general typea — the shaft furnace and the raerberaiory furnace. The first consiets of a vertical shaft or chimney, as in Fig. 62, while in the second the charge is kept separate from the fire.
THB SHAFT-FinUTACE
These are of two kinds, the wiud-fumace using natural draft as in Fig. 64 and the blast-furnace Fig. 65.
The Wind-funiace. — This is iised in the laboratory where it is desired to melt a charge in crucibles. When converted for use as a muffle furnace, the muffle ia inserted at the side, and is surrounded by the burning coke, which gives it a temperature high enough for melting and cupelling. As a variation from this, the flame from a soft coal fire, carried below the muffle, will sufficiently heat it. The wind-furnace is used in metallui^cal operations for melting down a product in plumbago or clay crucibles of 100 lb. capacity or over, these being embedded in the coke. Indeed crucible steel is still melted in this way. r,o.62.-Q,pol.Bl«,iFu,„». . The BUst-fmace -Fig. 66
shows the elements of a blast-furnace and the way it is chair?ed with coke and materials of the chai^ in alternate layers. It is seen that the charge and fuel are in intimate contact with one another so that not only is the chai^ effectually healed, but is brought within the reducing action of the fuel. 72
Dg,:,z..byG00glC
The Blast-Furnace 73
The Ct^Mla-fumace. — Ti^. 62 is a cylindrical blaet-fumace such as is used in foundries for the melting of pig-iron for castii^, and also in small plants for the treatment of copper ores. The cyhndrical shaft is wst«r- jacketed of steel plates. At the bottom are seen the two half drop-doors, which, when the furnace is running are swung up in position to cloee the bottom of the furnace. On them is laid the bottom lining of the firebrick. The crucible C, as high as the wind box E, is also Uned with 4^ in. of firebrick. The shaft above this point, to as high as the feed floor, is double, with a space of 4 in. between the iimer and the outer plates. Since this ^race is filled with water it is called a " water-jacket." The water here would soon reach the boiling point were it not for its circulation within, the cold water entering both at the top and bottom of the jacket by the supply pipe e, and the hot water leaving by the overflow pipe d. Thus, while the inner shell is in contact with the highly heated contents of the furnace, it is kept cool and is not melted or attacked by the slag, which indeed coats it. The lower part of the water>jacket is surroimded by the wind box E, as shown. There are six tuyere-openings through the water-jackets into the furnace, with openings opposite them in the wind-box as shown, these latter closed by covers having mica-covered peep-openings by which the condition of the tuyeres can be observed. In case of the stoppage of a tuyere, the cover is removed and a punch bar driven in to remove the obstruction. The base-plate rests upon four cast-iron columns, and sust^DS the entire structure. In the top T is the feed-door by which coke and charge are put into the furnace, the feed-floor being at the level (rf the sill of the feed-door as shown in Fig. 190. This figure also shows how the gases and smoke are carried away by a branching " downtake " to a dust-flue, or when desired let go above the roof of the blast-furnace building.
Other types of blast-furnaces are to be found in Figs. 152 and 155 for iron, in Fig. 195 for copper, and in Figs. 257, 25S and 261 for lead.
THE SEVERBERATORT PDRIfACB
This consists essentially of an enclosed fireplace or firebox at the rights hand side of the melting hearth B, Fig. 63 or as shown at a of Fig. 71, Here the fuel is burned, the products of combustion and the flame being drawn over the hearth to a chimney. The hearth is covered by an arched roof so as to reverberate or throw down the heat upon the charge placed on the hearth or sole of the furnace. This space is sometimes called the " laboratory," that is, the enclosure where the labor or work upon the charge is performed.
In case the furnace is designed for roasting ore, then the hearth is flat and level with the door aills, but when, as in reverberatory smelting, the contents of the furnace are melted, the hearth must be dish-shaped or bollowii^ and beneath the sill level so that the molten contents are retained.
, Fig. 63 gives a general idea of the appearance of a reverberatory roasting and melting furnace, a portion being broken away to show the hearth, side walls, and arched roof. It brings out clearly the way the furnace is " ironed," that is, tied together to resist the thrust of the arch and the expansion of the brick-work, as the furnace is heated. The ironing consists of upright buck-staves of railroad rails, one on each side of each door, tied across at top and bottom (the lower tie-rod beneath the floor level) by IJ-in. tie-roda. This is also shown in Fig. 71. Other buck-staves, set at the ends of the furnace and of the melting section or fuae-box B, are tied to reeiet longitudinal expansion. The furnace smoke passes away through an outlet port at the end of the furnace to the flue F and thence to the stack common to several furnaces. The fuse-box or hearth at B has two counterbalanced lifting side doors, as has also the fire-box, where is also to be seen the ash-pit beneath. The door openings DD have
Flo. 63. — Hand-rotating Reverberatory Furnace.
simple sheet-iron covers. Between them is seen a flat plate with an arched back rib to stiffen it. This takes the thrust of the arch between the doors. In F^. 213 is shown the heavy ironing needed for a large smelting reverberatory furnace, where, due to the high melting heat needed the expansion strains are great. The buck-staves then arc 6-and 8-in. I-beams with IJ-to 2-in. tie-rods. One must note here that as the furnace heat up the tie-rod nuts are slacked off a httle, and in cooling down they are tightened up — all to reduce severe strains. The open-hearth steel fiunace, Fig. 176, is similarly tied.
In the furnace, Fig. 63, the mfun hearth of the furnace is used for roasting. This completed, the charge is pushed down through a broad port into the fuse box B, there to be sintered or melted. When melted it is withdrawn by means of rabbles (hoes) at the side doors into wheelbarrows set beneath the door-plate.
The above description refers to a coal-fired furnace. There are three methods of firing as described under head of " Reverberatory matte smelting," i.e., direct or coal-fired, pulverized coal and oil firing. To this we may add producer-ga« firing, used for open-hearth furnaces in steel-making.
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Chapter Vii Combustion
PfilNCIPL&S OF coHBnsnoH
Combustion, as generally undersUxxl, may be defined as a vigorous chemical combination, attended with the production of hgbi and beat. To start combustion, the fuel must be (1) brought to Hts temperature of ignition; (2) it must be maintained at ttiis temperature; (3) a sufficient si^iply of air must be provided ; and (4) ttie products of combustion must be removed.
A jet (^ gas, burning as it issues from a tube, begins to take fire one or more inches from the tube, and continues to burn as rapidly as the molecules of the gas come in contact with those of the air. In an open-hearth furnace a current of heated ga& and one of air mingle gradually, ajid do not become fully mixed and inflamed imtil a few feet from the outlet ports. It is the same in a reverberatory furnace especially where there is but little more air than required for perfect combustion. The furnace, even if 100 ft. long, may be filled with flame from end to end, showing that gas at the distant end is still combining with air and burning. If, in the fire-box of such a furnace, we carry a thick coal fire, not less than 18 to 24 in. deep, the air passes through the fire freely. The flame is then shorter, since the hydro-carbon gas is speedily consumed. The long fiame is desirable where we wish to extend combustion through the furnace and not to pnv duce so intense a heat near the fire. If fuel-oil be fed into an oil-burner or injector, where it is broken up by a jet of steam or air into a fine spray, and then blown into a red-hot combustion chamber, it will bum Uke gas with an intense heat. Finely powdered coal projected in the same way, and with a sufficient supply of air (preferably preheated) burns rapidly, resembling gas, and furnishes abundant heat and high temperature.
Thus, to promote rapid combustion, the fuel must be in such form as to afford plenty of contact to the air. A piece of charcoal of large size bums readily, because, being porous, the air readily finds a way to penetrate it ; while a lump of anthracite, being dense, bums more slowly. Paper and kindling wood expose a lai^ surface to air, and hence ignite readily and bum rapidly. Paper in books, and' fabrics in bales, bum with difficulty. They may pass through fire only singed on the outside. Light lumber and boards bum readily, while heavy beams of wood resist a fire, with super* 75
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ficial charring. A thick layer of eawdust or fine coal, thrown on a fire may extinguish it. Hence, in firing up, such fine materials should be added sparingly, and used with lump coal or pieces of wood, to make passages or cavities, throu^ which the air may pass. Finally, as may be seen, a common error made in fire-buildii^ is that while an abundant supply of fuel may be present, insufiScient provision is made for the free passage of air through the fuel. A good draft and a sufficiently large exit-flue must be provided to carry away the products of combustion.
Flame is gas undergoing combustion. Soft coal and wood bum with a flame because the heat from burning distills, or drives out, the hydrocarbon gas which is formed. Anthracite, coke, or charcoal bums with little or no flame; while hydrogen bums with a non-luminous, though very hot fiame.
The temi>erature of ignition or of kindling varies according to tJie volatile constituents of the fuel. Thus bituminous coal, wood, and ordinary charcoal will kindle at toward 400° C. Anthracite and coke are hard to start, kindling at 700° C, a full cherry heat,
COHBVSTIOn IN TBB AIR AND m THE BLAST-FURNACE
Tlie Natural-draft Furnace. — As aji illustration of what takes place in a deep fire, let us consider a fire of glowing coke, Fig. 64. Here the air enters through the grate-bars, and, at the first instant, in contact with the glowing fuel, produces carbon dioxide,
with the development of a large amount of heat. Between 2 and 4 in. above the grate (with a clear fire) we may expect to find the highest temperature. This forms the zone 1, Fig. 64, and hence in a crucible furnace the bottom of the crucible should be set 4 in. above the grate to get the full efi'ect of the heat of the fire.
As we go upward, the CO2 in excess acta on the glowing carbon and dissolves or combines with it as follows :
This reaction is accompanied by the absorption of heat, and thus zone 2, Fig. 64, is cooler than the one below. In the zone 3, no reaction takes place, the fuel being simply heated by the ascending gases. A little air may pass along the walls, and issuing above the surface of the fuel, and mixing with the CO gas, burn a small portion of it to carbon dioxide with a blue flame thus:
This reaction also is a heat-producing one.
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Combustion 77
We finally gei, with a thick fire, a. mixture of gases of a compoeition much like the following:
N 70 per cent, CO 25 per cent, COa 2.5 per cent, 0 0.5 per cent and H 1,0 per cent.
The presence of the hydrogen is due to the deconipoeition of the moisture in the air. This mixture of gases can be made in a gas-producer (see Fig. 7), and for that reason it ia called producer-gas. Because of its content of CO it can be burned according to Equation (3), and used as a fuel for any purpose of heating. To burn the gas completely, and to get
Flo. 64.— Wind Furnace. Fia. 65.— Cupola Furnace.
the most heat from it, the thickness of the fire should not be greater than is shown in zone 1, Fig. 64.
Combustioii in the Cupola Furnace. — Fig. 65 represents a foundry-cupola charged with alternate layers of coke and pieces of pig iron. Here the object is to melt the iron, collecting it in a pool or bath at the bottom in the crucible of the furnace, shown in the illustration. Air is forced into the furnace, and fills all the voids, rising through the chaise chiefly in the passages or openii^ offering the least resistance. In an iron blast-fiimace (see Fig. 154) the rate of upward velocity approximates 6 ft. per second. As air meets the burning coke, combustion takes place according to Equation (1), producing a white heat. This action, in a cupola of 36 to 48 in. diajneter, extends upward about 3 ft. from the tuyeres, the upper limit of zone 1, Fig. 65. As the gases enter the zone 2, the CO2 just formed is decomposed by contact with the hot coke and forms CO, according to reaction (2), the change being nearly complete at the upper Umit of zone 2. In the upper zone no change takes place in the gases, and they impart their heat to the cold charge, which is
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being supplied as fast as the ore sinks below the required level. In this cupola, where the operation is one only of melting, to attain the greatest economy of fuel, the coke should be dense, the pieces lai^, and the blast abundant to supply plenty of air. Thus, burning the coke is deferred to the last, less CO is formed, and the combustion, performed largely in zone 1, is more nearly complete, developii^ the largest possible amount <^ heat.
From Equation (1) we find, that to bum one pound of carbon to COa, and thus with the greatest development of heat, there is needed 2.66 lb. o:^gen, or U.6 lb. air, since air contains 23 per cent oxygen by weight. At the sea-level 12,4 cu. ft. air weighs 1 lb. This makes 143.8 cu. ft. or, in round numbers, 150 cu. ft. air per pound of carbon. Ordinary coke contains 85 per cent carbon, thus requiring 122 cu. ft. air per pound of such coke. While in theory 12 lb. air should be sufficient per pound of coal, it has been found that excess is needed for complete combustion. For natural draft, using a thin fire, 18 to 24 lb. air has given the most satisfactory results, and where air is forced into a closed ash-pit, and through the fire-bed (undergrate blast), then 16 lb. air, or even less, is sufficient.
Figuring from Equation (2) in the same way, we find, per pound of carbon 1.33 lb. of oxy^n required, or of air 5.79 lb., equal to 71,8 cu. ft. Upon the basis of coke containing 85 per cent combustible matter, 61 cu. ft. air is required per pound of fuel when burned to CO.
Chimneys or Stacks. — In a furnace reaction not only is it neceBsary that the reaction elements be present in mutual contact, but that the products of the reaction be removed as fast as formed. Under conditions other than this the reaction ceases. A draft, therefore, must be provided, to cany away the waste gas, and to expel it into the atmosphere. This draft may be natural or forced. To insure obtaining a sufficient draft in a chimney, the gases must be delivered into the stack while hot. A temperature of 200° C. is ample for this, but since the work of most furnaces is done at a temperature higher than a red-heat, the excess may be utihzed to generate steam by conducting the gases through waste-heat boilers before entering the stack.
In a reverberatory furnace, used for smelting, the quantity and intensity of the heat depend upon the amount of coal burned per hour. This varies between 18 and 40 lb. per square foot of grate area and, to bum it completely, there will be needed 150 cu. ft. air per pound of coal consumed. In these furnaces the gases escape at temperatures between 300° and 1100° C, and move with a velocity of 12 to 20 ft. per second. For illustration, take the furnace Fig. 122, with a grate-area of 112 sq. ft., a consumption of 30 lb. of coal per square foot of grate-area per hour, needing 514,000 cu. ft. of free air. Allowing a temperature in this instance of 1000° C, and a draft velocity of 20 ft, per second at this high temperature, and knowing that these gases expand V273 of their volume for each degree
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Chimney Draft
above 0" C, we find the volume at 1000° C. to be — ^=5 — "4,7 times
the volume at 0° C, Assuming the temperature of the outside air to be 0° C, we shall have as the volume of hot gas per hour 2,368,800 cu. ft. At 20 ft. per second, or 72,000 ft. per hour, this will be an area of stack of
'ZrVTr -^ Bq. h. The actual area is 32 sq. ft.
Tbe total pull, or suction, that a chimney can produce, assuming it
to be filled with hot gases, is due simply to the ascensive force of tbe
gas measured by the difference between its weight and the weight of an
equal volume of the cold air outside. To maintain the velocity of the
gas in the stack, it has been found that a suction, or " pull," of 0.4 to
0.8 in. of water, as measured by a water-gauge, is needed. Taking a draft
of 0.6 in. in tbe above instance and fbdding 0.1 in. for friction in the chim-
62 5x0 7 ney, we have 0.7 in, water equal to — =3.647 lb. per sq, ft. A
cubic foot of air at 0° C. weighs 0.0807 lb. (12.4 cu. ft. per lb.). The gas
inside the stack has a specific gravity of 1.03 weight of air being unity,
«7X1
Hence we have the difference (0.0807-0.0177=0.063) as the ascensive
3.647 ine BiacK snouiu mereiore oc ~
From the above calculation it appears that the diBft^pressure varies with the temperature and height of the chimney. The velocity of the gas, or the amount of air passing through the fire per hour at a given t«mperature, varies as the square root of the height of the stack, in accordance with the equation V— v'2gA; g beii^ acceleration due to gravity and h head in feet. Thus a stack 100 ft. high would increase the velocity only 1.31 times more than our S8-ft. stack calculated above. The volume of gas increases directly with the temperature; while the velocity varies with the square root of this, hence there is a point of maximum discharge, at 273° C. At a higher temperature, while the velocity increases, the weight of the gas on the contrary diminishes.
Tehpbratdkb Of Cobcbustion
By this is meant the temperature of gases resulting from combustion under ordinary atmospheric pressure. We can calculate this when we know the calorific power of the fuel, and the total weight and mean specific heat of tbe resultant gases. The specific heat between 0° C. and the temperature <^ combustion increases as is shown in the diagram, Fig. 6S.
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Flame Temperature. — As an example of the use of the following tablei let us find the maximum temperature of combustion obtained in burning one pound of coke of 85 per cent carbon, using the theoretical amount of air or 9.86 lb. (since pure carbon requires 11.6 lb.), neglecting the loss of hefty in the adjoining walls of the furnace. Since by weight there is 77
Fia. 67. — Mahler Bomb Calorimeter.
I —
I
p-
I —
/
'
/
A
' 1
/
/
1*
n
/
J
' 1
2
0
0.
Fig. 68.— Specific Heat of Gasee.
per cent nitrogen in ^r, there will be, in the mixed gas resulting from combustion, 7,6 lb. nitrogen, and 3.11 lb. CO2 (0.85 lb. carbon, and ft X0,85-2.26 lb. oxygen). We assume a temperature of 2000° C, which is as near the desired one as we can judge. From Fig. 68 we find for 2000" the specific heat to be, for Na, 0.281 and for COa, 0.364. We then have:
Flame Temperature
The number of calories necessary to raise the entire gaseous product from 1 lb. of coke one degree will then be 3.266. The heat developed by the burning of the coke, being 6800 cal., the temperature of combustion is
TTUFi =2080° C, The specific heat of the gases at this temperature being
so nearly the same as at 2000° C, we can use it in this calculation. Were the difference great, we should have to use the specific heat at the exact temperature and calculate again upon that basis.
To proceed further, let us find the temperature of combu8ti<ni, or flame-temperature of carbon monoxide burning to carbon dioxide:
CO+0 =CO2=68,000 cal.
One pound of CO will produce 2440 cal., and will take 0.572 lb. oxygen or ^ = 2.48 lb. air containing 1.91 lb. nitrogen. There is also 1.57 lb. CO2 produced. Taking the specific heats from the table, Fig. 4.
1.91 lb. nitrogen @> 0,287=0. 548 cal.
1.57 lb. carbon dioxide # 0,364 = 0-572 "
2440 for each d^p«e of rise m temperature. Hence - -- « 2200° C. — the temperature of combustion. Again, calculating with the increased specific heat of the newly found temperature (2200°), we find 2112 to be the exao*. number of calories, nearly identical with that already found.
Fuel CombuBtion. — In writing equations the molecular weight is understood as in ordinary chemical equations.
(1) C+02 = 97,000 may also be written
with a comma to indicate th&t the different molecules, bo separated, unite to form CO2.
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S2 Combustion
If oxygen btima in presence of an excess of highly heated carbon, theo carbon monoxide is formed, and this may be written
which indicates that by the combination of the solid carbon with the gaseous oxygen 29,000 calories have been formed. Likewise this may be written
Since the 12-lb. carbon gives 29,000 calories, we have, as the heat evolved by the bummg of 1 lb. of carbon, 2440 calories. If we bum the 00 thus formed with a sufficient amount of air, we have
or, as also written, C0,0 = 68,000 calories. Were this written C,03 then we would have 97,000 calories (Equation (2)). Equation (5) may again be written
(6) (^0+0 = C02. 29,000 97,000 = 68,000
This means that before this reaction can take place, the CO must be broken up according to the reaction,
or again
the minus 8^ meaning that in thi? reaction as much heat has been absorbed in the breaking up as was earlier evolved in Equation (3), in which these elements united.
In Equation (6) the CO havii^ been decom[K)sed into C and 0, they are forced tOAinite with the 0 to form COa, evolving 97,000 calories. The net result or the algebraic sum of the two reactions is thus 68,000 calories as given in Equation (6).
Tempera ores of Combustion. — The temperature of an incandescent body may be judged by the eye or by an optical pyrometer according to the color scale herewith;
Lowest visible red 470° C.
DuUred 550 to 625°
Cherry red 700°
Light cherry red 860°
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Flame Temperature
Onnge 900° C.
YeUow 950 to 1000°
Light yellow 1050'
White 1150'
Dulling white 1500 to 1600°
These ccdors apply both to flame and to a heated body.
Fig. 69.— Wail Type Indicating Pyrometer.
Indicjitiiig Pyrometer. — Fig. 69 is a Le Chatelier pyrometer composed of a thermo-couple of a platinum and a platinum rhodium wire placed at the end of a protective tube of porcelain which is threat into the fire or against the heated object. The two wiren are continued as leads to the terminals of a galvanometer graduated to indicate the temperature.
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CHAPTER VIII METALLURGICAL THERMO-CHEMISTRT BCETHODS OP DETGRHIfCmG THBRHIC-VALUBS
Deflnitioii of Metallurgical Tlieniio-CheiniBtry. — This pertains to questions relating to the heat evolved or absorbed when elementary substances or their compounds combine in metallur^cal operations. By having an intimate knowledge of these reactions we can utihze fuel and control roasting, smelting, or converting operations to the best advantage.
Thus when cupeUing rich lead in an Enghsh cupelling furnace. Fig. 181A, the air passing over the hot molten bath is seen to aid in maintaining its molten condition, energetically uniting with the lead to form litharge, and so evolving heiA.
When wood or other fuel is kindled, the properly supplied and regulated draft of air maintains the combustion of the fuel with evolution of much heat.
A mixture of quartz-bearit^ ore, with fluxes in suitable proportion, the whole brought to a whit« heat, will melt tc^ether, the heat being intensified in so doing.
On the other hand, when steam is passed through a glowing coke fire, it in decomposed, forming hydrogen and carbon monoxide, but absorbing heat and cooling the fire.
Units of Measurement — ^The amount of heat generated as the result of the combination of elementary substances is given in Table I. In Table II is given the amount of heat evolved as the result of the combina ■ tion of certain bases (oxides of the metals), aa given in Table I with silica.
The unit of measurement used is the heat required to raise a unit weight of water one degree. Thus, one gram of water raised in temperature one degree Cent^rade is called a small calorie (cal.) or gram-calorie. One kilogram of water raised 1° C. is called a large calorie (Cal.) or kilogram-calorie. One pound of water raised 1° Fahrenheit is called a British thermal unit (B.t.u.) and it is but 0.252 of the kilogram-calorie. The gram-calorie is used in small-scale laboratory operations. For ordinary work the large calorie is preferred.
In this book and in the calculations which follow we shall use the pound-calorie, this being the heat required to raise 1 lb. of water 1° C. When we use the word " calorie " the pound calorie will be understood. The kilogram-calorie ia 2.2 times greater than the pound-calorie.
General Prim^les. — 1. The amount of heat needed to decompose a
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Calorimetry 85
compound into Us constituents is equal to that evolved when that compound is f oimed from those constituentB. When a reaction takea place by which heat la absorbed, as in Eiquation (8), it is called " endothermic." On the other hand, when heat is evolved in a reaction, as in Equations (1), (3), and (5), it is said to be " exothermic."
2. The heat evolved in a chemical process is the same, iriiether it takes place directly or in several steps. Thus in Equation (3) the carbon is burned to CO with the evolution of 29,000 calories. The CO thus formed, when burned with additional oxygen, as in Equation (5), gives 68,000 calories, and the sum of these two is 07,000 cat., the same as if the carbon had been burned to C02 as per Equation (1)
In comparing reactions (1) and (3), it may be said that in presence of an excess of oxygen, reaction (1) would take place rather than reaction (3). This is in accordance with the law of Berthelot, namely :
3. Every reaction vtuch takes place independently of the addititm of energy frcnn without the system, tends to form Hm combination which is acconqtanied by the greatest evolution of heat.
Calorimetry. — To determine accurately the heats of combustion of fuels, or the heat of formation of compounds, the Mahler bomb-calorieeter. Fig. 67, is much used. It consists of a steel shell or bomb, marked B, shown also on an enlai^ed scale at the right-hand upper comer of the illustration. The bomb, holding about a pint and weighing 9 lb., is shown to be closed by a screw-cap, having a stop-cock threaded connection x by which it may be connected by a flexible pipe to a cylinder 0, which contains compressed oicygen gas. Within the bomb is suspended a capsule c in which is placed a gram of the substance to be tested. The cap is then t^tly screwed on, and oxygen gas under a pressure of 300 lb. per square inch is allowed to enter. The shell is next placed in the calorimeter D, which contains a known weight of water. The thermometer T is set in place, and the stirrer or agitator iS is set in motion to bring the whole apparatus to the same temperatiu^. The calorimeter D is placed within a larger vessel A covered with a thick layer of felt and provided with a thermometer (not shown). The vessel A serves also to support the bracket G from which the stirrer is suspended. The temperature of the calorimeter having been noted, the charge is ignited by a coil of the plaV inum wire F. The resultant rise in temperature is noted by the thermometer T. The total heat developed, with certain corrections, is calculated from the weight of the calorimeter water, and from the rise in temperature. In those oases where the heats of formation of oxides or of sihcates are desired, the net result is accomplished by respectively oxidizing or melting them in the bomb with a known weight of a well-determined fuel. The Ttumber of calories evolved is the algebraic sum of those of the desired reaction and that of the fuel.
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80 Metallurgical Thermochemistry
HEATS OF FORMATION OF THE ELEMENTS Following are the mcdecular weights and the heats of formaticm (^ some of the better-known chemical compounds. From these may be estimated the heat developed in various reactions:
Table I
Heat Of Formation Of Chemical Elements
Moltcolu Weight..
Formula.
Co
s,o,
c,o,
s,o,
Hi,0
Micon and PluKphorM
Si,0,
P,0.
Oxidet of the MetaU
MgO
Fe,,0,
Ba,0
Fe.,0.
Ca,0
Sb„0,
A1,,0,
Sb,,0.
Na,,0
A9„0,
K,,0
Aa,.0,
Mn,0
Cu,0
Mn,0,
Cu„0
Zd,0
Pb,0
Fe,0
Sulpkidtt
Pb,0, ftheMelah
Ba,S
Cu„S
Ca,S
Ou,S
Zn.S
Pb,S
Fe,S
Sb,,S, ftheMeiaU
B«,C,0.
Mg,C,0,
Cft,C,0.
da in DUule Sotutim
Zn,S,0.
Na.,S,0,
Fe,S,0.
Ca,S,0,
Fe,,S.,0,
Mg,S,0.
H„8.0,
Al„S.,0,
Cu,S,0.
Mn,S,0,
Anudgams of GM and iStltuT XHk,Au Hg in exceed 197 2,580 | XHg.Ag Hg ii
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Heats Of Formation
TABLE II - Htatt of Formation of the Silicate*
Farmmtion HmI.
PeO,SiO,
80
2FeO,SiO,
MD0,SiO*
41
1S3+ 60-213
69
2CaO,SiO"
J«8+ 60=228
SiO,,SiO,
AI,0,,2@iO,
67
86
2HAAl.O.,28iO,
Li,0,SiO,
Narf>,SiO,
S6+102-158
3
15
3CaO,AI,0,
11
MnO 1 0+CaO 11.4 +
MgO 2 7+Al,0, 9.2+
CuO 0 4+8 0 4%
Fe057.6; CaO 12.0; SiO,30.4%
The heat of formation of silicates, if we were to start from the elements, as in Table I, commonly amounts to from 2000 to 4000 calories per pound of the compound thus formed; but, when the metals and silicon have become ondised, as commonly occurring, moat of the heat of formation has developed. If then such oxides and sihca are brought to melting temperature they combine with a farther development of heat, as given in Table II, which varies from almost nothing up to 720 calories per pound of the silicate formed, or an average of 145 calories. Thus, in Table I, the heat of formation of FeO is given at 65,700 calories and of silica at 180,000, or tc^ther ^5,700 calories, being 1100 per pound. But when the slag FeO, Si02 is formed only 10,600 calories is developed, equal to 80 calories per pound of the slag.
HEAT EVOLVBD AS THE RESULT OF ROASTINO In the reaction C+ 02 = COa (see Table I, Carbon, Hydrogen, and Sulphur) one equivalent, or 12 lb. of carbon ie completely burned by combining with 32 lb. of oxygen, forming 44 lb. of cfirbon-dJoude and evolving 97,000 calories. Dividing this by 12 we have 8080 calories as the result of the burning of 1 tb. of carbon.
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Chapter Ix Roasting
By roasting we mean the preliminary treatment of ores by fire at temperatures below their melting point in order to improve their condition for subsequent reduction or extraction. Because of the expense the operation is avoided where possible.
We may classify these operations into: (1) Calcination or kiln roasting; (2) oxidizing roasting; (3) chloridizing roasting: (4) sulphatizingroastii^; (5) sinter roasting.
Caldnation or Kjln-roaBting. — Carbonate ores, as of iron or zinc, are charged in lump form together with some coal or wood into a vertical kiln. The heat expels the cont^ned moisture and carbon dioxide of the ore. In the'case of some Mesabi iron ore, containing 12 t0'15 per cent of moisture, freight is saved by kiln drying and the ore is made more porous and accessible to the action of reducing gases in the subsequent smelting. The grade of the ore is raised and a better price is obtained for it; if a siderite or iron carbonate, the weight is still further reduced.
Oxidizing Roasting. — This is for the purpose of expelhng the moisture, of burning off the sulphur, and in the case of certain refractory ores, of removing their contained arsenic or tellurium. Thus, these are freed from impurities which would interfere in subsequent smelting or leaching. The ore is heated to burning temperature with free access of air, and gives a porous product, easily penetrated by reducing gases or by solutions. As a treatment for leaching, roasting destroys colloids, so that the roasted ore is more easily leached. The product of an oxidizing roast is often called calcines, though this is better applied to the product of calcination.
Chloridizing Roasting. — It is performed in a reverberatory furnace, common salt being added at a certain stage of the roasting to change the ore to an easily leached chloride. Advantage is taken here of the principle of mass-action, whereby the nascent chlorine reacts on the oxidized metalliferous products.
SulidiAtizing Roasting. — This is a partial oxidizing roast, some of the sulphur being expelled, and the remainder changed to SOa- Thus a sulphate of the metal is formed, which is soluble and therefore can be extracted by water-leaching.
Sinter Roasting. — Sulphide ore spread out on a travelii^; grate is heated
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Chemistry Of Roasting 89
by a flame to bumii^ temperature. Air is then drawn through it by aid of a suction fan, resulting in a vig(H-ous oxidation, expulsion of the aulphur, and a partial sintering together of the ore particles. Not only is the ore well roasted, but the dust loss io the subsequent blast-fumace smeltii^ is greatly lessened. Iron ore in too fine condition for blast-fiunace smeltii^ has been mixed with 6 to 10 per cent of its weight of fine coal to sinter or agglomerate it. In the past much fine ore has not been so treated, but has been fed in the crude state so that the blast-fumace has made a high flue-dust loss.
Chemistry Of Roasting
Chemistiy of Ozidizing Roasting. — To do good roasting, we should have (1) heat sufficient to start the burning; (2) a final heat sufficient to drive off the last portion of the srulphur; (3) preferably 46 lbs. of air per pound of sulphur, (4) an extensive surface exposed to the air; (5) frequent stirring in order to present to the air fresh surfaces for rousting.
Let us take the case of an ore with a silicious gangue, containing the sulphides, pyrite, chalcopyrite, blende, and g^ena. This is dropped upon the hearth at the hopper end of the furnace, Fig. 63, and then spread out. Here the temperature, 350° C. is sufficient to expel moisture and start the reaction of combustion. In ten to fifteen minutes burning begins, as evinced by a blue fiickerii^ sulphur Same that plays over the surface of the charge. The pyrite is thus decomposed.
The first, loosely held equivfdent of sulphur is easily expelled and unites with the air, with the evolution of 3220 pound-calories per pound of sulphur burned. The FeS now remaining, together with the other sulphides, begins to oxidize. The FeS and CuS is most easily oxidized, while the ZnS and PbS are the slowest in parting with their sulphur. Beginning then with the FeS we have
or in words, the iron sulphide becomes oxidized to ferrous oxide with the formation of S02, the reaction being exothermic, and yielding 113,600-;- 32='3550pound calories per pound of sulphur burned. Cupric sulphide of the chalcopyrite reacts according to the formula:
or per pound of sulphur, 98,000+32 = 3030 Cal. The blende under the . action of air and heat is affected in the same way.
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or per pound of Bulphur present, 3420 Cal. Galena roasts according to the reaction:
which gives off 3250 Cal. per pound of sulphur.
It will be noticed tiiat the heat evolved per pound of sulphur is much the same in each case, and hence the sulphide highest in sulphur yields the most heat. These reactions, especially of blende and galena, are gradual during the roasting period. The air acts chiefly on the exposed surfaces and hence roasting is hastened by stirring the charge. That FeS which is near the surface has an excess of air, and in presence of silica which acts by catalysis, it becomes oxidized thus:
(6) 3FeS + 110 = 2SO2 + FejOs + FeS04, 3X23,800 2X71,000 199,400 235,600 = +505,600
or per pound of sulphur, 5260 Cal., indicating an energetic exothermic reaction. Of the products of the reaction the sulphur dioxide is carried away by the draft. When the two are stirred together, the FeaOa is acted on by FeS as follows:
(7) FeS + lOFeaOs = 7Fe30* .+ SO2.
23,800 10X199,400 7X265,800 71,000= +86,200
FeaO* is of a black color; when the ore is roasted to excess the resultant product is the red FeaOg, often an undesirable, red-colored product.
As the charge is moved toward the firebox, the iron sulphate produced as in (6) begs to decompose at a temperature of 590° C. and in presence of cupric oxide reacts as follows:
(8) FeSO* + CuO - FeO + OUSO4 235,600 37,200 66,400 182,600= -23800
That is, the 80g given off by FeSOi while in nascent condition is taken by the CuO to form its sulphate, the reaction being an endothermic one. Such of the FeSOi as is not decomposed by the cupric oxide is brokeil up by the heat alone as follows:
also an endothermic reaction.
At a slightly greater heat (655° C.) the cupric sulphate, formed but a
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Chemistry Of Roastino 91
short time previously, begins to decompose and at a dull red heat the decomposition of cupro-cupric sulphate begins. These reactions are complete at a cherry red heat (850° C.) up to this temperature. These are the reactions of a sulphatuig roast.
At 850° C. the zinc and lead oxides, reacting on the copper Bulphate now decomposing, begin to be changed to sulphates thus:
(11) . PbO + CuSO* = PbS04 + CuO. 51,000 192,600 216,200 37,200= -19,800
These last four reactions are endothermic and instead of aiding the roasting absorb heat as the result of the reactions. Fortunately they take place in the hotter part of the furnace.
As the charge is moved nearer the fire the above just-formed sulphates decompose, the zinc sulphate more readily than the lead sulphate, and SOb escaping.
At 1050° C. (an orange heat) copper oxide is decomposed into cuprous oxide, and ferric oxide, long some oxygen becomes FesO^.
At this stage the ore begins to fuse if it contains lead, but with Httle lead it sUghtly agglomerates, with much it fuses. The chai^, now no longer porous, ceases to roast, in fact it is hard to roast such an ore well. On the other hand a zinc ore, free from lead, can and should be brought to a high finishing heat to decompose zinc sulphate and to eliminate sulphur.
To decompose completely a lead-bearing zinciferous ore for further treatment in a Uast-fumace, the foUowing procedure is successful. After roasting to the point of fusion, the ore is removed to a reverberatory melting furnace. Here silicious ore is added and the whole melted at a high heat. The silica reacts on the lead and zinc sulphates thus:
That is, the sulphur is eliminated as a sulphuric anhydride, leaving a sulphur-free sihcate of both metals. In the blast-furnace the zinc silicate enters the slag as such, while in the presence of fuel, the lead is reduced and recovered.
At 590° C. the iron sulphates, formed at a lower temperature, begin to decompose.
At 655° C. copper sulphates, formed at a lower heat, begin to decompose.
At 705° C, already-formed cupro-cupric sulphate (CuSO*) begins to
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At 850° C. copper sulphates are entirely decomposed, and when steam is present the msjdmum amount of soluble sulphate (AgSO^) is formed.
At 1100° C. ferric oxide (FejOa) is decomposed to the next lower oxide FeaO*.
In oxidizing roasting it has been found that with 2 per cent SOa by volume, or 4.4 per cent by weight in the escaping gases, roasting is active. This corresponds to 46 lb. (570 cu. ft, at sea level) per pound of sulphur driven off. Calculating this for a 16-ft. MacDougall roaster treating 40 tons of ore in twenty-four hours, and roasting it from 35 per cent S down to 7 per cent Hulphur, we have an elimination of approximately 0.25 lb. sulphur per second. This needs 142 cu. ft. of free air, equal to 284 cu. ft. of the temperature of 273° C, that of maximum chimney discharge. For a velocity of 20 ft. per second in the stack as a maximum this would require an area of 14.2 sq,, ft. or a diameter in a round stack of 4 ft, 3 in. With an excess of air above that just specified, the hearth tends to cool off; with less, roasting proceeds more slowly, bo that at 4 .4 per cent SOa in the escaping gases, the roasting goes actively, at 8 per cent very slowly and at 9 per cent it ceases altogether.
The larger the charge and the greater its thickness, the longer is the time needed to complete the roast. A few grams are roasted in a half-hour in the muffle, and in twenty hours in a reverberatory furnace, while it takes weeks to roast ore in the pile.
We may note the temperatures of the reactions that occur in the reverberatory furnace as follows:
At 150° C. the odor, due to the volatilization of some of the loosely held or first equivalent of sulphur, can be detected.
At 350° C. the sulphur of the sulphides {particularly of pyrite) begin? to bum with a blue flame.
Heap Roastiog has the advantage that it can be used at the first installation of a small smelting plant in a new district, where it is aimed to avoid investment in an expensive roasting plant. It requires only the necessary site, the method is a simple one, and the results are satisfactory. On the other hand in a laigc plant, where from 10,000 to 50,000 tons of ore or matte are in process of treatment, heap-roasting may cause the locking-up of several hundred thousands of dollars in the heaps.
Th« Chemistry of He^> Roasting. — The heat generated in the burning pile volatilizes sulphur from chalcopyrite, and where there is insufficient air some of it escapes from the top of the pile as elemental sulphur. The remainder, uniting itself according to the equation: S-|-02=S0b, leaves
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Heap Roasting 03
as sulphur dioxide. In coutact with heated ore it ia further changed to milphur trioxide, S02+0 = S03.
Air haB access to the exterior of the lumps, and the reaction on the iron-copper sulphides is as follows:
Site of He^. — For the heap the leaching sites should be nearly flat, but with drainage to one border, layered or coated with clayey slimes and sprinkled with fuel-oil for tightness. The yard for a large site should have three service railroad tracks, two for the green or turoasted ore, run along outside the piles 170 ft. apart, the roast-ore tract midway between these. This leaves room for two rows of roast heaps each 60 ft. wide, 100 ft. long and 8 ft. high, to hold 2500 tone.
Heap Building. — In building a pile or heap the foundation is laid, usually of deadwood, to a depth of 12 to 18 in., the surface of the wood being roughly leveled. At intervals of 10 It. are left flues which are filled with small wood, to be ignited in order that the fire may penetrate rapidly to the interior of the heap and produce a more uniform combustion. Coarse ore, amounting to about two-thirds of the whole, is then piled on the wood, followed by a layer of medium-spied ore, and lastly by fines which cover the top and sloping sides of the heap. A supply of fines is kept close by, bo that, wherever the pile ia burning too fast in any given spot, it can be more deeply covered, and where it seems dead, the layer can be opened up to encourage the fire to that spot.
After lighting, the wood bums out in about sixty hours, leaving the ore in vigorous combustion. The pile will bum for three or four months with occasional regtilation of the draft as above described.
During the roasting, the outer portions of the piles become reddish in color, due to the oxidation of the iron, and a httle sulphur condenses on the surface, but is later "
driven off. The raw ore may be specified as averaging 23 per cent sulphur, and this is reduced to 10 or 12 per cent. Fig. 31 shows a cross-section of a lump of well-roasted ore contuning copper.
The copper sulj^ide is not so easily decompo»ed, due to the greater aiEnity ot copper for ^^ ,„ _„^^ I.„„p sulphur as compared with iron. It fuses and O^
accumulates as a layer beneath the oxidized crust.
As this crust gets thicker so also does the layer, until finally we find the lump made up of iron oxide with a center of copper sulphides of a bronze color, and, given time, even this sulphide becomes oxidized. This final condition is shown in Fig. 70.
A certain amount of sintering or fusion always takes place and ports of
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the heaps have to be blasted loose. When the roasting is completed the ore is dug out by steam shovel, loaded upcm cars standing on the center track and taken to the blast-fumaceB.
Since there is no protection from the weather during roasting, and Eonce soluble sulphates of copper are formed during the operation, a portion of the valuable metal is leached out by rain and snow-water. This loes is estimated at 1) to 2 per cent.
. The cost of roasting at Ducktown, Temi., at the first of the century is said to be 42 centa per toD, but at a low wage. Peters gives a cost for fuel, labor, and supplies 48.5 cents per ton with conmion labor computed at $1.50 per day. Heap roasting may be done by contract to advantage. At the United Verde, Jerome, Ariz., 75 cents per ton was the contract price. The Canadian Copper Co., in 1916, roasted its ore on a large scale for 50 cents per ton.
Heap Roasting of Matte. — Matte can be well roasted in lump form, but unlike ore, it requires two or more burnings. After the first firing, in spite of care, matte shows but little the change it has undergone. At the second burning, using a larger quantity of wood, the result of the first burning begins to show. A lai^ portion of the twice-burned material is found to be light in weight and porous, and to contain no unbumed core. In fact the thoroughness of the roast may be judged by feeling of the lumps with the hand. If well-roasted lumps are broken, they uo longer show the raw core at the center.
The bed of wood can be prepared for matte as for ore, but the pile is smaller, being only 12 ft. square by 6 ft. deep, with a single chinmey at the center. The broken matte, with the raw fines spread over it, is covered with the finer portion of roasted material. The burning of the heap lasta eleven daya, and when ended, it is taken down, and the imperfectly roasted part made into a new pile, and the roasted matte sent to the furnace. It is a good plan in constructing the new pile to introduce one or two layers of chips or bark, for a reducing effect upon impurities like arsenic, and for producing a more uniform heat throughout the pile. Finally, after this burning, a large portion suitable for use can be sorted out and the part still incompletely burned can go to the next heap.
Roasting Of Orbs In Pdlvbrizbd Condition
This work is done in single or multiple-hearth furnaces. The ore is spread out upon the hearth or floor in about a 4-in. layer, exposed to the action of flame and air.
If not already tine enough, it is crushed so fine that the particles at the end of the time given for roasting show no unbumed core or center. An ore, mainly iron pyrite, decrepitates in roasting, hence is,fipe enough
ROASTING IN tTttlNACES 95
if of two-or threeHmesh size. Many ores and matte need crushing to four-or six-^nesh sise. An ore of blende or galena is compact and when to be roasted (especially when " dead roasted " so that no sulphur is left) had better be ground to lO-mesh size. Where the ore is finely groimd for subsequent leaching this may be done before roasting. It is a good plan, however, to grind to a coarse size for roasting, and to regrind the roasted product as fine as desired for after treatment.
In the various furnaces advantage is taken of the heat developed by the burning of the sulphides, especially in the compact multiple-hearth roasters. If the percentage of sulphiuis high, this is often enough to supply the required heat (after combustion has once been started) without the aid of extraneotu fuel. Thus in the MacDougall roaster, after the furnace and ore has been sufficiently heated, a content of 25 to 30 per cent Bul[Aur ensures the continuance of roasting.
The various mechanical roasters treat ore cheaply, but for ores containing much lead, which a^lomerate or sinter, they do not work well. With a slight accession of heat above the normal, caused by lack of care in firing, the ore is liable to agglomerate, and eventually to stick to the hearth, stopping the movement of the rabbles. When this becomes serious a stout flat bar of iron, attached to one of the rabble arms in place of a rabble blade, may plow up these accretions, and by setting it in different positions on the arm the hearth may be finally cleared. The device has not proved entirely successful. In the hand reverberatory roaster the hearth is accessible, and when the accumulation builds upon the hearth it may be removed by aid of cutter^bars and a hammer. If, however, it is sufficient to rough-roast such an ore, reducing the sulphur content, no more than from 10 to 13 per cent, then such a^lomeration need not be feared. Hand roasters work well upon ores that need a high finishing heat suited to breaking up or decomposition of the sulphates, as in the roasting of zinc ores or of galena. The objection to such roasting is that it is costly.
Thb Lohg-Hearth Reverberatory Roaster
In this furnace the charge is put in and removed at intervals. These furnaces may be distinguished from the reverberatory furnace used for melting by the relatively small grate area and by the fact that the hearth is flat and at the level of the door sills. The hearth may be 10 ft. wide by 36 ft. long, divided as shown on the plan. Fig. 71, into three hearths with a drop of 2 in. between. (Large furnaces are built 70 ft. long with five hearths 14 ft. wide.) The length of a hearth for a reverberatory roaster should accord with the percentage of sulphur Uiat the are contains, and in consequence the heat developed by it in roast- "'i- Without the aid of the heat developed as the result of the burning of
Section 0-D
Via. 71.— Reverberatory Rooating-tunwce {eectionB).
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Mechanical Roasting Furnaces 97
the Bulphur,*the fiie would not maintain sufficient heat to roast ore 25 ft. from the fire-4>ndge. An ore containing 10 per cent sulphur can be roasted to good advantage in a furnace having a mngle hearth of 15 ft.; whm 15 per cent sulphur is present we may add another hearth, bringping the length to, say, 30 ft.; a 20 per cent ore would work rapidly in a three-hearth fiu-Qace ; an ore of 29 to 33 per cent would do well on a four-or five-hearth furnace.
To furnish draft for a stack or chimney (see the plan Fig. 71), 28 in. diameter inside by 65 ft. high will be sufficient.
Operation. — Into the thoroughly hot furnace (the slide of the charge-hopper being withdrawn) a charge of 2000 lb. pours in a conical heap on the first hearth and ia there spread by a man on each side using a paddle. (This has an iron pipe-handle 12 ft. long with a blade 6 in. wide by 18 in. loDg.) Here the ore remains for four hours being stirred every half-hour with a rabble, a hoe having a blade of 6 by 10 in.
By means of the paddles it is then moved down and spread out on the second hearth, while the first hearth receives a fresh charge fnHn the hopper.
Again at the end of four hours both chaises are moved toward the fire, and another charge dropped on hearth No. 1. Thus all the hearths become covered with ore. At the expiration of the four-hour period the first chai^ now on the last hearth is withdrawn through a square discharge hole seen near the fire-bridge in the figure. It drops into a wheelbarrow set beneath. This ore has thus been under the action of the fire for twelve hours, and for the charge specified we compute an output of 6 tons daily of raw ore or as much as 5 tons of roasted ore or calcine.
ICBCHAinCALLT OPERATED ROASTinO FDRHACES
These may be classified as follows:
(a) The revolving cylinder furnace with the axis horizontal or inclined toward the dischai^ end. An example of such a furnace is the Brtickner cylinder roaster, having a horizontal axis, the cylinder 7 ft. diameter by 25 ft. or more long. The ore is roasted in batchen or charges. It is charged and discharged through manholes. A charge of 40 tons may take three days to roast. The Oxland, the White-Howell, and the Argall are examples of cylinder furnaces with the axis inclined, causing the ore gradually to travel from the feed end to the lower or discharge end, as the furnace revolves. The Oxland and the Wliite-Howell are amgle-cylinder furnaces. In the Argall four are united in one.
(b) The mechanically rabbled rererberatory furnaces having a continuous feed and discbarge. Of these the Brown-O'Harra, the Ropp, the Edwards, the Merttm, the Wethey, and the Hegeler have strai^t hori-
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zontal hearths and ao are called stnughtr-line furnaces. The first four have single-hearths, the fifth has two superimposed hearths and the Hegeler is a multiple-hearth furnace.
Another variation of horiztrntal hearth furnace has the hearth curved
as in the Brown horseshoe furnace, or circular, as in the Pierce tiuxet furnace. The ore in either case having made the circuit of the hearth is discharged. The rakes or rabbles are but part of the time in the furnace in order that they may have time to cool. The Brown horseshoe has but a single hearth while the Pierce turret may be single or double.
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The White-Howell Furnace 99
A type of fum&ce once used in the chloridizmg roasting of silver ores waa the Stetefeldt where the ground ore was showered down a shaft.
The Vhite-Howell Cylinder or Furnace. — Fig. 72 is a longitudinal elevation of this furnace. It consists of a cylinder, 50 in. inside diameter by 34 in. loQg, set at an inclination of 2\ per cent supported cm friction-rollers carried on the drinng shaft. At one end is the firebox, at the other a dust'.chainber which connects by a flue to the stack. The hotter end of the cylinder, near the firebox, is of larger diameter, to permit of its being lined with brick, thus leaving the cylinder of uniform interior diameter throughout. Projecting, longitudinal, firebrick ledges, set spirally, raise the ore and shower it back through the flame as the cylinder revolves, so as to roast it more rapidly. The unlined part for the same reason b furnished with longitudinal, cast-iron, projecting shelves. Ore is fed at the flue-end, by means of ascrew-feed (see Fig. 267), and when dropped into
Pia. 73. — Cylinder Dryer.
the revolving cylinder, travels along, dischai^g at the firebox end. Just before it reaches the firebox it passes out from the cylinder to a brick chamber below, and is withdrawn from that when cool. The furnace makes much flue-dust. It is used chiefly for chloridizing roasting, upon ores containing but little sulphur, and has a capacity of 50 tons per twenty-four hours for low-sulphur ores.
For an ore-drier a furnace quite like this is employed, as shown in Fig. 73.
The Edwards Roasting Fumace.^This is a single-hearth reverberatory furnace with hearth dimensions 57 ft. long by 6 ft. wide. Fig. 74, in plan, shows a portion at the firebox end, the feeding mechanism and the cooling floor in section. The elevation shows the side, constructed like a plate-iron beam, the stirring mechanism and the conveyor for transferring the roasted ore to the cooling-pit. Fig. 75 is a transverse section of the hearth showing the details of the stirring mechanism. The slope of the furnace can be changed a httie by tilting. This regulates the rate of travel of the
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The Edwards Roasting Furnace 101
ore thiou^ the furnace; but for a given kind of ore, tliia slope, once determined, is not again changed. The furnace has a slope of ) in. per foot toward the dischai^ or firebox end. The stirring and propulsion of the charge are effected by means of rabbles fixed to vertical shafte, as shown in the elevation of Fig. 75, and in the plan of Fig. 74. 'nieae rabble-shafts make one revolution in sixty to ninety eeconda. The rabbles at the firebox end are watetH»oled, and thia is found especially necesaaiy wbere a high finishing heat is needed. The blades or plows of the rabbles can be easily replaced through the doors adjacent to them. The figure indicates the hearth as broken away, at the dischai^ end, to show two of the rabbles in plan. The last rabble sweeps the roasted ore into the discbarge shoot, and the push-conveyor
then moves the ore to the coohng- „ _ „ , . „ , „
■^ FTfl_ i_ .. t ^L F""- 75.— EdwardB Routing Furnace
pit. The bottom of the conveyor (iVimvene Section).
trough is furnished with shoes, by
means of which the ore can be dropped at any desired point on the coolii^-floor. The ore is fed to the furnace from the feed-hopper, by an endless-screw conveyor which discbat^^ into a feed-opening in the roc^ of the furnace, l^ie smoke is earned off by a Sue. The ftunace takes 1 H.P. to operate, and baa a daily capacity of 25 tons on sulphide ore of 30 to 35 per cent sulphur. The roasted ore contains 3 to 8 per cent of sulphur. The moving parts are durable, and the furnace has proved efficient in practice. I^arge installations, of the duplex type with a double instead of a single row of rabbles, and of hearth-dimensions 120 by 12 ft., have been built for a daily capacity of 60 tons. These furnaces do not have the tilting hearth.
Besides these we find circular revolving-hearth roasters, as the Bnmton and the Spirlet. The Brunton furnace is iised in the roasting of arsenic-bearing fiue-dust. The Spirlet is used in blende roasting; the hearth revolves and the rakes or blades are fixed in the roof above.
The Multiple-hearth Furnace Type. — ^A group of mechanical roasters of the circular, multiple-hearth type are to-day most, used. These are the MacDougall, the Wedge, and the Herreshoff. Of them we will fully describe the MacDougall and the Wedge roasters.
The UacDougnll Roasting Furnace. — ^There are several kinds of furnaces of this type. Among these are the Herreshoff and the Wedge. The MacDougall furnace as manufactured by the AUis-Chalmeis Co. is shown
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in sectional elevation in Fig. 77, It is a vertical, cylindrical furnace, 18 ft.
diameter, with ax arched hearths, over which travel rabbleB which stir
and move the ore gradually toward the drop-openings through the floor
of each hearth, situated alternately at the center and at the periphery. A
central shaft is provided, carrying six radial rabble-arms (three of these
rabble-blades set at an angle
on the arm.
The rabble-blades on the even-numbered hearths are so set as to push the ore in a spiral path toward the periphery; the odd-numbered ones toward the center. The ore, fed continuously into the fiu-nace from a cylindrical hopper shown above and at the right. Fig. 76, drops upon the upper hearth near its outer edge. The rabble-blades of that hearth stir and move the ore gradually toward the central drop-openii^ where it falls to hearth No. 2. The rabbles of this hearth again stir and move it to the outer
„ „ , _ drop-openings, through which
Fia. 76.— Sut-hearth MacDougaU Roasting .^ , ,, , . ".,_ vt o -m_
Furnace it falls to hearth No. 3. The
ore advances by this means
until it reaches the lower hearth, where an opening at the periphery gives
it erit to a receiving-hopper, shown beneath the hearth, from which it is
drawn into a car as required.
A high-sulphide ore roasts by its own heat when the furnace is in fuK operation. The ore fills the hearth to the level of the blades, and is spread out evenly by them. On the upper hearth, as the ore moves toward the central opening, it becomes dry and hot, and when dropped upon hearth No, 2, begins roasting. On hearth No. 3, the ore roasts freely, emitting sparks and forming sulphates. On hearth No. 4 no sparks are seen, and the ore has attained its highest temperature. On hearth No. 5 the ore looks less bright; and on No. 6, especially at the dischai^, it has bocome cooler.
The air for oxidation is admitted by side doors, mostly those of the lower
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THE MacDOUGALL ROASTING FURNACE 103
hearths. The gas and dust, pasaing up through the drop-openingB, are drawn thrcH^h the horizontal main flue. In starting, the furnace is heated to the kindhng temperature of the ore which, if rich in sulphur, bums by its own heat, without the aid of fuel. If the sulphur content is low, additional heat is supplied by one or more external flreplaces, near the bottom of the furnace.
To protect the rabble-anus from the intense heat they, and Ukewise the central shaft, are water-cooled. The cooling-water is forced down the 94n. hollow, central shaft in a 3-in. pipe to a point near the bottom, and out to the ends of the arms in 1-in. jnpes. It then returns up the annular space between the 3-in. [npe and the hollow shaft, and discharges at the top through two spouts into a launder. The furnace is 18} ft. high by 18 ft. diameter, and has a total hearth-aiea of 1600 sq. ft. The structure is supported on columns to give room below for the hopper and the car into which the roasted ore is discharged. The shell, made of |-in. plate-steel is lined with 9 in. of brick-work. The rabble-arms consume 1) to 2 H.P. and make one revolution in IJ minutes.
A furnace treats, in twenty-four hours, 65 tons of sulphide ore of 35 per cent sulphur, reducing it to 7 per cent. About 4 per cent flue-dust ia made ; and the ore itself contains more ferric oxide, and is Ughter and more porous than if treated in a hand-reverberatory roaster. The cost of roasting such ore is approximately 35 cents per ton, which is the lowest figure thus far known for any furnace. The compact form of the furnace reduces radiation to a minimnni and permits roasting with Uttle or no fuel. Taking ca[)acity into consideration, the furnace is one of moderate price, and one that costs little to keep in repair.
Of the two revolving-hearth furnaces, the Holthoff and the Raymond, the latter has some popularity for the preUminary roasting of ores for blast or pot-roasting, the powdered ore bemg showered down a vertical shaft or tower and coming in contact with an upward flame from a firebox. An objection to its use is that much flue-dust is made.
Pig.76 is a perspective view of the enclosed type of MacDougall roasting furnace, where, in order to permit a firebox below the lowest hearth the driving mechanism has been transferred above the furnace. The firebox enables the hearths, and especially the lower one, to be heated. By its use the ore may be roasted to a lower percentage in sulphur.
Id Fig. 77 we give 'a plan and sectional elevation of a MacDougall roaster. At the center drop holes, a plate on the rabble arms hold up the ore close to the opening so that the gases do not pass upward there. The outer drop^oles are similarly sealed. Thus the gases must pass upward by the gas passageways, and are thus free from flue dust.
The Wedge Roasting Furnace. — Fig. 7S is a sectional elevation of a seven-hearth furnace as constructed for oxidizing roasting. If the ore
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Pta. 77. — MacDoug&l Roasting Furnaces (sectioDs).
The Wedge Roasting Furnace 106
contains moisture, this b dried out upon the top, called the drier-hearth. The central hollow shaft A is covered outside with tiles to protect it from the furnace heat. It carries at each hearth two opposite rabble-arms which are water-cooled by small feed and discharge pipes leading down from above. Individual pipes from the water pan E pass down the shaft to the
Fic. 78. — Sectional Elevation of Wedge Roaster,
rabble arms. The hearths are enclosed in a steel shell 22} ft. diameter. The outlet for the sulphur-bearing gases is at F.
The driving pulley P through a train of spur gears communicates motion to a bevel pinion on the end of a horizontal shaft, and this meshes into the master gear G, at the foot of the hollow shaft A . This same gear has a turned raceway at its edge supported by roller that cany the whole weight of the shaft and the sixteen rabble arms.
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Fio. 79. — Elevation of Wedge Roaster.
Fig, 80.— Section of RoasUr Plant.
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The Brunton Roasting Furnace 107
The OTe, fed at the out«r edge of the top or drier hearth, works down under the central plate upon hearth No. 1. It passes across the hearth to
Fig. 81. — BrunWn Roa-tting Fumsfe.
the outer drop-holes of hearth No. 2. On hearth No. 3 the ore has drop- ' holes other than the central opening where the gaeee rise so there is less
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flue dust made and ho on. The lower hearth has a peripheral drop-hole for the discharge of the calcine. F^. 78.
The method of installation of a roaster building contaming multiple-hearth roasters is shown in Fig. SO. Cmshed ore from the sampling and grindii^ building is carried to its top story by an incline belt conveyor and 18 delivered to the roaster feed-hopper by means of an automatic tripper aa the horizontal 20-in. conveyor belt.
The Bronton Roasting Furnace. — In ¥ig. 81 we show a plan and elevation of this furnace, also sections of the firebox. This has a revolving hearth and three sets of rabbles, being iron blades paeeing through the roof. These, set at an angle, stir and move the ore. which is fed through the roof at the center to the discharge chute at the right. The hearth is carried on a cast-iron frame fixed to a vertical shaft and is driven by a lai^ woimwheel and a worm on the horizontal driving shaft. There are two coalfireboxee kk, the flames entering at the side of the hearth and the mingled fire gases and arsenical fumes escaping at flue m.
Roasting Of Matte
The term " roasting " is apphed also to a method of treating copper matte in a reverberatory furnace in large pieces, upon which an oxidizing flame is i^owed to play. Such masses slowly melt and are acted on by the air, whereby a part of the material becomes oxidized or roasted sufficiently for the next operation. As compared with ordinary roasting this is slow, and the method is one but little used.
Copper-bearing matte to be subjected to an ordinary oxidizing roast must be crushed at least to 4-mesh zize. Matte from the silver-lead smelting to be roasted in a reverberatory furnace of the kind shown in Fig, 63, needs a different treatment from that given to ore. This kind of matte contains but 20 per cent sulphur, and does not take fire like pyrite ore, but must have a high finishing heat to expel the sulphur. Such matte is considered well roasted when it contains 4 per cent sulphur. Ores low in lead can easily be roasted to 2 to 3 per cent sulphur, while galena, when roasted, still contains 5 to 6 per cent when drawn from the furnace. Like matte, galena starts burning slowly, and must be roasted slowly, for rapid heating causes it to ^ter and thus stops further roasting. Typical leady matte contwis metals and sulphur as shown in the sub-joined table.
The roasted low-grade matte contains 23 per cent oxygen. This explains why it does not lose weight in roasting. Pyrite ores of 20 to 30 per cent sulphur, on the contrary, easily lose 15 per cent in weight.
Losses in Roasting. — Such loss depends U[>on the extreme to which the roasting is carried ae well as upon the nature of the ore. When ore is so
Roasting Costs
Per Cut.
1U».
Cu
roasted that it vb not sintered at the final high temperature, the lead lost averages 2.5 per cent, but no loss of silver occurs. When the tempera^ ture is carried higher, and the ore is ag^omerated, the loss is sightly h^her. When fused it may reach 15 to 20 per cent of the lead and 2 to 5 per cent of the silver. Of the gold little is lost in oxidizing roasting.
CAPACiTT OP fuhhaces and cost of roasting
These depend upon the surface exposed to the oxidizing influences and upon the quantity of sulphur contained in the ore in hand reverberatory roasting. Silicious ore, containing } to 3) per cent sulphur, requires 13 to 15 sq. ft. of hearth-area per ton of ore roasted per twenty-four hours. Matte contaming 20 to 25 per cent sulphur, when it is necessary to reduce the sulphur content to 4 per cent, needs 45 sq. ft. hearth-area; copper sulphide ore, roasted to 7 per cent in preparation for smelting, requires 33 to 35 sq. ft. For roasting iron-sulphide concentrate, which carries 35 to 45 per cent sulphur, down to 3 to 10 per cent sulphur, 55 to 60 sq. ft. hearth-area is needed.
Roasting Costs. — In 1910 to 1913 ore-rotating in heaps, at Jerome, Arizona, cost 80 cents per ton , including general expense. Ore-roasting in stalls cost 60 cents per ton. For reverberatory roasting, in long, hand-rabbled furnaces the lowest price attainable on copper ores was $1.50, with an average of $1.81 per ton. For roasting lead-bearing ores, $1.75 is a moderate coat, and from this the cost, when all items are included, may rise to $2.25 per ton. The AUen-O'Haira automatic furnace, having two straight hearths each 94 by 9 ft., and resembling the Wethey furnace, roasts 45 to 50 tons daily at a cost ot 78 cents per ton. The Wethey furnace, of the type having four hearths, each 65 by 10 ft., the roasting proceeding on all the hearths, roasts 90 tons daUy to 5 to 6 per cent sulphur, at a cost of 98 cents per ton. The 16-ft. MacDougall furnace (Herreshoff type), having five hearths, 14J ft. diameter, and a total area of 830 sq. ft., roasts 33 to 35 tons daily to 7 per cent sulphur at a coat of 50 cents per ton. The Briickner roasting cylinder, 8^ ft. diameter by 22 ft. long, takes a chai^ of 20 tons (10 tons daily), and in forty-eight hours roasts it to 4 per cent sulphur at a cost of 80 cats per ton.
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It will be noticed that the low cost of roasting in some of these fumacea is due to their needing no fuel after coming into full operation. To obtain this effect such furnaces have several he&rths, and are compact. On account of this compactness they lose but little heat by radiation.
The above roasting costa were for the period 1910 to 1913, but these figures must be doubled for present conditions.
Blast- Or Pot-Roastiko Of Orss
Both lead and copper ores are treated by blast-or pot-roasting, though the method was at first intended for lead-bearing ores, especially for galena. I have already mentioned the dif&culty of roasting galena by the old method, in the reverberatory furnace; but by pot^-roasting, it can be 8o treated as to remove most of its sulphur, with less loss by volatilization.
Treatment ot Galena. — ^By the Huntington-Heberlein process, called also the " H and H process," the galena^bearing ore is given an incomplete, rather rapid roast, to reduce the amount of sulphur to 12 to 14 per cent. The product from the roaster is mixed with a certain proportion of limestone and siliciouB ore, wet down, and charged into a hemispherical casL-iron pot 8^ ft. diameter by 4 ft. deep, having a capacity of 8 to 10 tons as shown in Fig. 82. Within the pot, and forming a false-bottom, is placed a circular arched plate perforated with |-in. holes to admit air to the charge under pressure. Upon the false-bottom is scattered a wheelbarrow-load of ashes, then a carload (one tun) of hot ore from the roaster. On this is dumped 8 tons of charge wet to about 6 per cent moisture. Air, under the pressure of a few ounces, is admitted beneath the false-bottom, and comii^ up through the hot ore, it produces a burning-temperature and starts the cranbustiou of the charge. The heat gradually ascending to the top, the char^ becomes red-hot, and SOz and SO3 escape. At the end ctf the roasting, which lasts sometimes sixteen hours, there remains only 3 to 5 per cent sulphur if the charge ia properly burned. The pot is now inverted to discharge the contents, and this falls out in an ^glomerated, red-hot mass. It is broken to a size suited to subsequent treatment in the blast-furnace.
The Dwight-Lloyd Machine. — Blast-roasting in pots has several disadvantages: the ore is exposed for a long time to the hot gases and this leads to a loss of metal ; the process is intermittent ; the charge needs constant attention either in chai^rng, discharging or blowii^; the amount of fine is apt to be considerable and this must be re-treated; the ore is not evenly sintered; finally, it is expensive to break up the sintered mass, whether by hand or by power. These disadvantages appear to be overcome by the Dwight-Lloyd sintering process, especially by the use of their endless-chain machine, 28 ft. total length, illustrated in Fig. 83. As
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The Dwight-Lloyd Blast-Roaster
compared with pot-roasting, and particularly with roasting furnaces, this machine occupies but little room.
The endless-chain carries a train of pallets. Each pallet is in fact a perforated grate having two edges uptiuTied 4i in. The joints between the
Id at ti>[>, with I''hofe Fia. 82. — Details of Construction of Blast-roasting Pot.
pallets and between the pallets and the suction-box are close and fit snugly by pUuied edges. At the ends of the suction-box a planed dead-plate, over which the pallets glide, serves to make the joint tight there. After the pallets leave the suction-box their four wheels transfer their weight to the rails. At the delivery-end, where a car is stationed, the pallets are slightly
raised, fracturing the Bintered coke, then drop one by one, striking the pallet next below on the circuit, while the sintered ore is jarred off into a car. The ore is fed to the machine through a pug-mill where it is wet to 4 per cent moisture by a spray, and falls to a feed-hopper set 4 in. above the peJlets,* thus ensuring a layer of ore of that depth. It moves along at the rate of 12 to 20 in. per minute. As it reaches the ignition-box it b set afire by an ignition-fiimace burning coal, or otherwise gasoline is burned, using a series of Bunsen burners supplied by air imder pressure. The suction-box 12 ft. 6 in. long by 30 in. wide is connected to an exhaust-fan at a vacuum of 6 oz. As the ore passes the burner it is ignited, and the air, sucked down through the layer of ore, continues the burning, which is completed by the time the ore reaches the end of the suction-box. There
Fill, 83.— Straight-line Dwight-Lloyd Blast-roa«Ung t'urnace.
results a product which retains 3 to 5 per cent of sulphur only. By the time the ore layer has reached the discharge-end it is aoUd. The mouth of the feed hopper is set to give the required depth of material, and the layer of ore is smoothed by a stiff brush 30 in. wide, then by a roller, all tending to level and compact the ore to ensure even sintering and roasting. Care must be taken that the layer of ore is of uniform density and that there is no segregating of it in the feed-hopper; otherwise the coarser part of the layer bums rapidly while the denser part does not get enough air.
The mixture or charge may consist of lead concentrate, fine oxidized and silicious ores, together with flue dust. Its composition may be quite variable. A satisfactory mixture will carry 35 per cent SiOa, 18 per cent S, and 20 per cent Pb; another one, 16 per cent SiOa, 18 per cent Fe, 15 per cent S, and 30 per cent Pb.
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Sinter-Roasting Reactions 113
Tbe machiiie will treat 40 tons in twenty-four hours at a cost of 75 cents per ton. One to two horse-power is needed to drive it.
Reactions m Sinter-roastfaig of Copper-bearing Sulphides. — ^When a mixture of iron and copper sulphide is sinter-roasted, desulphurization proceeds rapidly if the ore be wet and silica be added ; otherwise it proceeds slowly. For these reactions we have:
(15) 3FeS-|-4HaO = Fe304-|-3HaS+2H,
When air is drawn through the charge both hydrt^n and HjS bum, but FeaO* reacting on FeS gives FeO as follows:
This reaction is exothermic and at a high temperature with silica would form ferrous sihcate, again producing heat. Indeed, in action the formation of this, with the consequent sintering, can be seen spreading as the burning proceeds. The cost of blast-roasting in pots has been given at $1.44 to $1.80 per ton.
Triple Roastiho
While blast-or pot-roastmg is less used than formerly, in one case, viz., at East Helena, Mont., zincky Cceur d'Alene concentrates high in lead are roasted in three stages, A Godfrey revolving-hearth roaster brings it down 13 per cent sulphur, .^ain crushed, it passes to a Dwight-LIoyd ffliiter machine which reduces the sulphur to 8 per cent. This product, if smelted, would produce too much matter, therefore, after again crushing, it is blast-roasted m Huntington-Heberlein pots, and yields a final material that carries but 2 per cent of sulphur — so low b this element in fact that there is no production of matte in smelting it.
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Chapter X
CONCENTRATION OF ORES AS A StJBSIDIART OI^RATION IN METALLURGT
In mills treating ore containing a heavier part, such as sulphides, gold or amalgam from the plates, it is customaiy to catch, such heavy products:
1. By gravity concentration in which the heavy particles of the pulp are caught on blanketed or canvas-covered surfaces;
1. The Blanket or Canvas Table. — ^This conosts of a floor, sloping like an amalgamating plate, 1§ in. to 3 in. per foot. It is smoothly covered
FlQ. 84.— Wilfley Concentrating Table.
with blankets or canvas, and along its upper edge the mill-pulp ia flowed in an even, thin layer from a head launder to distribute over the table. The lighter particles flow down the slope, the heavy portion sinking into the interstices of the fabric. After a while the till-flow is switched to another table or floor. Fresh water or solution is then run on the loaded table which sweeps away the settled pulp,* leaving the heavy particles in the-interstices of the blanket or canvas.
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The Wilfley Concentrator 115
2. The Concentntiiig Table. — ^There are different varieties of Uiese, such as the Frue Vanner or the WilSey table. The mill-flow is received in a continuous stream at the head launder of the machine and delivers two products, a head product or concentrate, and a tail product freed from the heavy portion.
The Wlifley Table.— This may be taken as a type of table, which like the Overstrom, the Deister, or the Burchart is much in use in mills. I^ig. 84 is a front view of a No. 6 Wilfley table having a deck, a surface partly riffled, partly smooth, so transversely inclined that a sheet or film of water,
Fio. 86.— Wilfley Table in Action.
composed of feed-water conveyii^ ore-pulp and the washing or dressing water may be caused to flow across it. This deck, moimted on bearings, has an endwise reciprocating motion of ) to J in., imparted by a mechanic^ device caDed the " head motion." This head motion causes a separation of the heavy and light grains into layers by its agitation and jerking action, throwing them toward the head end, the lightest-grained being washed down the slope and dropping over the edge at the tailings side. This separation is well shown in Fig. 85. The heavy grains travel along in the riffles and are forced more toward the head, escaping there and at the bead-end-corner. E^ach kind is caught in its own launder.
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116 Concentration Of Ores As A 8Ubsidlary Operation
Oil Flotation
Flotation is a method of concentratmg whereby a finely ground water pulp, containing sulphides, as of lead, copper, zinc, or magnetite, tt^ther with a gangue of quartz or other equivalent mineral, is treated by addition of oil (or other chemicals), and violently agitated. The oil or other chemicals act to produce a film upon the sulphide particles, causing them to float as a froth upon the surface of the water, while the gangue sinks to the bottom. The froth is removed, Bettled, and filtered, yielding a product containing the valuable sulphides. This method possesses the advantage that it is effective upon slimes hard or impossible to concentrate by gravity methods.
Most plants use a mixture of various oils and the kind and quantity should be worked out for each particular case. Pine oil is quite commonly
Elevation Cross-Section
Fia. 86. — The Minerals Separation Machine.
used, also wood and coal-tar creosote. Creosol or cresyhc acid w used in small proportions with coal-tar. Other substances are fuel-oil, oleic acid, and eucalyptus oil.
Minerals Separation Flotation Machine. — Referring to the two views of a double machine. Fig. 86, the feed enters the first agitation box at the motor end of the machine, thence it passes to a second box, throilgh an opening in the partition wall, as shown in the transverse section.
From the second box it descends to the first spitzkasten where the froth, which rises to the surface, flows over the front edge into a narrow deep launder. The remaining pulp passes through a pipe {the inlet controlled by a valve) to the third agitating box. From this box the pulp passes to the second spitzkasten, and so on through the mac^hine until it reaches the fourteenth. The discharge from the fourteenth spitzkasten leaves the machine as tailings. The catch-launder is so divided that the first four to seven spitskastens make frothed concentrates, the remainder make middlmgs that are returned to the system. Some of the pulp is overflowed from the last three together with the froth. For affecting the agi-
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Flotation Machines 117
tation, e&ch of the sixteen vertical shafts nmnmg at 250 R.F.M. carries a four-bladed impeller. At Anaconda 6 to 8 lb. of 60 per cent sulphuric acid, 2 to 3 lb. of kerosene sludge acid, and ) to 1 lb. of wood creosote ia used per ton of flotation feed. Part of the creosote is added ahead of the tube-mill, where the pulp is groimd, the remainder with the acids at the first agitation box. The pulp ia heated to 70° F. by blowing live steam into it at the head of the machine. A machine will treat 175 tons daily of slimed pulp. Other flotation machines much used are the Callow and the Jann^.
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Part Ii Gold
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Chapter Xi
GOLD ORES AND CLASSIFICATION FOR BmUNO
Gold occuiB in nature, both in the native etate and combined with tellurium.
Hative gold occute in vetn^natter disseminated in grains or particles of various sises, and it is found not only in quartz veins, but in veins or lodes containing hematite, iron-pyrite, arsenical-pyrite, blende, and galena. Id pyrite it occurs not only in the substance of the crystals, but as ^ms on the surface of these crystals. It is frequently accompanied by silver. When gold-bearing veins have become disint^rated and swept away into alluvial deposits, the particles of gold, where released, are found in the Band and gravel of the beds, the pebbles and boulders themselves (which have come from the country rock), being in general barren of gold. Gold occurring in this way is called alluvial gold, and is recovered by methods of hydraulic mining or dredging, which belong to mining engineering rather than to metallui^. We shall consider, therefore, the treatment of gold ore.
Gold TeUuiides.— In South Dakota, at Cripple Creek, Colo., in Western Australia and elsewhere is to be found gold combined with tellurium as calaverite, AuTeg (containing 41.4 per cent Au and 57.3 per cent Te); also gold and silver combined with tellurium aa sylvanite (AuAg)Tea, and as petzite (AggTe, AuaTe).
Physical PropertleB of Gold. — ^This, the only yellow metal, has a specific gravity of 19.3 and is the most malleable and ductile of all metals so that an ounce of it will cover 160 sq. ft. It is softer than silver, harder than tin, and has a tenacity of 14,000 lb. per square inch, with a 30.8 per cent elongation. It melts at 1063" 0. and begins to volatilize at 1100^ C, so that by the time it reaches 1250° C. it volatilizes four times as fast.
These metals are commonly found t<^ther in ores, and a metallui^cal method suited for the extraction of one is often as well suited to the recovery of the other. The recovered metals, alloyed with one another, and forming a gold-silver bar, are shipped to the mint or to the refiner in that form for final parting into their constituent metals. Both gold and silver are won from their ores by milling or by ameltii^ methods, the former being the commoner way.
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'122 Gold Ores And Classification For Milling
Valuation of Gold and Silver. — The amount of gold and silver in or^ is expressed in ounces, in pennywe^hta, in gruns or in kilograms per ton. When in the form of bars or ingots the value is ^ven in the percentage or in the fineness of the respective metals. In English-speaking America silver and gold in ores and by-products are designated in ounces per ton or in dollars. Thus 0.05 oz. equals SI per ton, valxiing gold at S20 per ounce. Its exact mint value is, however, $20.67 per ounce. In the British Empire, except Canada, notably , in South Africa and Australia, the pennyweight is preferred, a convenient designation, ^ce it gives the value,xery close^to SI or fouTyShillings English money. In lAtin America gold is expressed in grams, silver in kilograms per metric ton. The contents of a gold-silver bar is given in fineness, i.e., in parts per thousand. Thus such a bar, upon assay, may be 750 fine in gold, 150 fine in silver, the remaining 100 parts being mainly copper. United States gold and silver coins are 900 fine in gold and silver respectively, the 100 parts remaining being copper. Sterling gold or silver is 926 fine.
Referring to the milling of gold ores, that Ib, to ores in which gold is the dominant metal, these may be treated directly, or as supplemented by concentration as follows:
Chlorination;
Cyaniding;
Amalgamation and cyaniding;
Amalgamation, concentration, and cyaniding the tailings;
Where concentrates are made these are further treated by smelting, oi at the mill by cyanidii^ them. They may be produced either by gravity (KHtcentraticm or by flotatiw.
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Chapter Xii Amalgamation
Plato Amalgamation. — ^Thia is the time-honored method for the recovery of gold from " free-milling " ores and as an important preliminary proceaa vHh gold ores, which though not free^nilling, contain a considerable proportion of readily amal^mable gold. By free-milling ores we mean those in which the gold occurs native and can be caught by amalgamation. However, the cyanide process is used after amalgamation where a portion of ihe gi^d is not recoverable by amalgamation.
Ores N eedfaig Amalgamation. — Free-milling oxidized gold ores in which the gold is shown upon panning, and where many of the particles are too coarse for solution by cyaoiding, are the ones for this method. At the Manhattan Big Four mill, where the ore is a soft calcareous schist with laminations containing calcite and quartz, but no sulphides, the fine gold, oceuiring in the laminations, is stamp-and tube-milled as just above outlined. It is thus seen that when the ore has been stamped coarse, the pulp is not run over the amalgamating plates until finely ground, so in this mill we find the plates not in their customary place below the mortar, but in a separate building, where they receive special attention, and are safe from pUfering.
Whether inside amalgamation is the most suitable is often discussed; but for certain ores it is the correct method.
Gold sometimes occurs in certun ores with a metallic appearance, but is mostly brown and lusterless. In this supposedly allotropic form it fails to attach itself to the amalgamated plate.
There is a consideraUe difference of opinion among metallurgists regarding the extent to which amalgamation should be iised in a goldmill. ScHne think with very fine grinding to bring all the gold (and any sflver) to the degree of subdivision necessary for rapid solution, that amalgamation may be dispensed with, and that by " all-sliming " the gold can be extracted by cyaniding.
The two extremes in the plate area provided for amalgamation prior to cyanidation are represented by the Rand and the Homestake. Figures frcHu thirteen Rand plants, for the year 1913, show the recovery by amalgamation on mill feed, approximating |6, to be 61.4 per cent the plate area ranging from 0.25 to 2.0 sq. ft. per ton milled per day, and the mercury 123
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conmunption about O.I oz. per ton milled. The tendency on the Rand has been to eliminate amalgamation directly after the stamp in favor of amalgamation after tube-milling, and at the eame time, reduce the plate area to leas than a half. In no case has the reduced plate area caused serious decrease in recovery by amalgamation, some instances showing no decrease, while, naturally, there is a decided saving in mercury-loss, attendance, and capital cost. On the other hand, amalgamation at Homestake is carried on both inside and outside of the stamp mortars and after regnnding. Of the total plate area of approximately 11 sq. ft. per ton milled, only 1.60 per cent is used after regrinding. In 1914, on $4.11 mill feed, the Hcnnestake recovery by amalgamation at the stamps was 69.0 per cent, and after regrinding, 0.8 per cent; the total of 69.8 per cent being effected with a loss of mercury of 0.13 oz., and at a total cost of 2.45 cents per ton milled.
Amalgamation, as practiced at the average plant, costs between 5 and 10 cents, the plate area approximating 2 sq. ft. per ton milled. The recent trend has been to relegate to this step the recovery of such coarse gold as may greatly retard, or entirely escape subsequent cyanide treatment. With amalgamation, crushing is nearly always done in water, although a few plants have been able to maintain their plates in fair condition when crushing in cyanide solution. The extremely low cost of recovering gold by amalgamation demands that this step be given weighty consideration. An \m\usual method of amalgamation at the Nipissing high-^ade mill is to be noted. A chaise of 3 tons of 2500-O2, silver ore (containing 39 per cent arsenic, 9 per cent cobalt, 6 per cent nickel), 4 tons of mercury and 1.5 tons of KCN solution is tube-milled, with oompressed air fed into the tube-mill for some ten hours, at the end of which period amalgamation has recovered some 97 per cent of the ^ver, the residues, freed from amalgam and mercury, then going to the cyanide plant. The mercury consumption is 20 lb, per ton of ore.
STAMP-MILLING WITH PLATE AMALGAMATION The ore is crushed to a size of 20-me8h or finer, using 6 to 8 tons of water per ton of ore, and running the ore pulp over plates of at least 4) ft. wide by 6 ft. long per battery or often much longer. The heavier gold soon touches the plate where it becomes incorporated with the amalgamated surface. About every shift the battery is stopped for a few minutes and the gold'bearing amalgam is scraped from the plates, and treated to obtara the gold. Amalgamated plates may also be placed inside the mortar; particles of gold adhere to these plates when driven against them by the splash of the pulp. Gold particles also fall to the bottom of the mortar to be caught by mercury there. The gold caught inside the mortar is recovered in the monthly clean-up. From time to time about 1.5 oz. menniiy per ounce of gold in the ore is added in the mortar as the crushing proceeds.
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The Stamp Battery 125
Piste ama^amatton is used for gold ores only.
The tailing, if barren, is run to waste. If it contains gold-bearing pyrite this may be caught on concentrating tables, see Fig. 84, to be eent away for smelting or to be 8[>ecially treated on the spot by cyaniding. The tailing, generally gold-bearing, would be cyanided.
The Stamp-battery. — Fig. 87 is a viev of a ten-stamp battery of wood
Fw. 87.^P(TBpwtive View of Ten-stamp Battery.
construdjon. The parts are thus designated. At A, the mortar block or foundation; B, the mud sills. C, the cross sills; D, the side posts; F and G, the buck-«takes; H and /, the lower and upper guide-timbers respectively. The foregoing parts constitute the battery frame, J, J are the cast-iron mortars as shown in section in Fig. 88. At /C is a wire cloth, or, as here shown, a slotted screen. At L is the die, resting on the sole of the mortar. The stamp consists of a stem having on it tappet P, by which it is lifted and
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at its lower end a boss N, carryiixg ite Bboe-M. these four parta being the total weight. In Fig. 88 this weight is 1250 lb. R is the cam shaft carrying ten cams U (better shown in Fig. 88), with its driving pulley V at one end. At K ia seen one of the amalgamated plates, already referred to, the other being omitted to show the screen and chuck-block belongiiig to the first five stamps.
In Fig. 87 the mortar blocks are heavy posts, set on end, and extending down to the solid rock or concrete foundation- Instead of these, ccmcrete blocks. Fig. 88, are preferred, the mortars being held down by long foundation bolts with l-in. sheet rubber and i in. shee1>-lead interpoeed to give an even bearing.
Fig. 88 is a sectional elevation of a ten-etamp battery unit, grouped as shown in Fig. 87, into two sets of five stamps, Ore from the feed bin is fed to the stamps by a suspended Challenge ore-feeder, which can be run out of the way when making repairs to the stamps. At £i is shown a horizontal lever depressed at each stroke by a collar on one of the stamp and by a vertical link and lever working a ratchet feed. As the circular feed-plate slowly revolves against a fixed scraper, the ore is scraped off to fall into the mortar. As the ore accumulates under the stamps, the stroke shortens and with it the feed, while as the mortar empties the stroke is increasesd and with it the feed. The ore from the lip of the feeder falls into the mortar, there to be stamped, and when sufficiently fine, the resultant ore pulp is driven through the screen at the front by the splash caused by the dropping stamps, to flow.over the apron plates (not here shown). The guide timbers cany the lower and upper caslHron guides for each stamp stem. At X is a finger-bar, one to each stamp, by which the stamps are " bung up " when, for any reason, a battery is to be stopped without stopping the other battery.
Operation ci the Stamp-battery. — The feed is regulated so aa to cause the stamps to strike with a sharp, hard blow, but with little of the rebound that would occur with a thin layer of ore.
Mercury Fed to die Battery. — This will average 1.5 oz. per obnce of gold caught. Added, a little at a time inside the mortar, it works out in part upon the apron plates. For the amount to be used the mill-man is guided by the appearance of the plates. If they are bard it means too little mercury or " quick " ; if the mercury shows on them in streaks or patches then too much is being fed. Mercury should be free from base metals that would cause it to "sicken " into coated globules, which would be swept away with the pulp. It is better that it contain a little gold.*
• The lo98 of mercury may axgregate 0.5 on. per ton of ore treated. It m»y be lort by flouring, indicated by a wbite appearajice,-and due to excessive agitation in the ur, which breaks it into particlee bo fine that they unite no more. Mercury may be lost by " Bickeninft," shown by a black appearance and due to the presence of base metals as already explained.
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The Stamp Battery
Fio. 88.— Scctiomd Elevation of Ten-stamp Batteiy.
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Dressing ttie Plates. — This is done three or four times daily, and takes about fifteen minutes. To do this feeding is stopped bo that the ore may work out of the mortar, and the stamps are hung up. The amalgam on the plates, perhaps a cupful, is removed with a rubber-edged scraper. If the surface of the plate is hard a little mercury is sprinkled on it. Where the plate has become tarnished by a verdigris coating this may be removed by salammoniac applied with a scrubbing brush. In a few
Fia. 89.— Clean-up Pan.
minutes this should be washed off, and potassium cyanide, then mercury rubbed on, and the plate washed clean. The stamps are now started, and feeding is resumed.
Apron-plates are set at a grade of ) to 1 J in. per foot, the steepest grade where sulphides occur in the pulp. When properly flowing over the plate, the pulp travels down in a series of ripples, thus bringing the gold in contact with it.
A mercury-trap receives the flowing pulp at the foot of the plate.
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The Clean-Up Pan 129
Any non-adherent particles of amalgam are here caught. The amalgam is occasionally removed by the plug-hole on the bottom erf the trap. Its overflow passes generally to the concentratii^ tables.
The Cleon-up.— This occurs once or twice a month. Let us take the case of a 40-stamp mill. Two batteries are himg up. The screens, inside plates, and dies are taken out. The ore in the mortars, two or three bucketsful, are taken and fed to the next batteries. The inside plates are scraped and dressed, and all is replaced, and the batteries again started. The next two are treated in the same way as well as the last ones, whose mortar contains an accumulattcm from them all.
The Clean-up Pan, Fig. 89, 3 ft. diameter and making 12 to 15 R.P.M.,
Fio. 90.— Vertical Retort.
is used for grinding the sand, pyrite, fragments of iron, etc., the accumulation of the last batteries. The charge, perhaps 300 lb., is wet^round to a fine mud with the addition of 50 lb. of mercury during the shift. The pulp is diluted with water and the muddy portion is run off by a side plug. The residual mercury and amalgam with some mud is withdrawn through the lowest plug-hole, panned in a gold-pan by hand, treated with nitric acid, and well washed until clean. The residual amalgam is strained through .canvas to remove the excess mercury. Gold amalgam this treated contains 35 to 45 per cent gold; the filtered mercury still retains 0.5 per cent gold.
Retmting. — In the smaller gold mills the amalgam is retorted as shown in Fig, 90. The retort, filled two-thirds full of amalgam and the cover luted
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and clamped, is placed in a wind-Jumace, there supported on a cast-iron rest. A water-cooled pipe leads out from the cover, its lower end dipping in a tub of water. The retort is graduaUy heated, the mercury vapor cconea over, and is condensed in drops in the water-cooled pipe and collect in the tub below. Tlie retort is kept at a distilling temperature for one or two hours, then heated to redness to expel the last of the mercuiy. This mercury is used again.
The remdue, taken from the retort, is porous and is from 600 to 900 fine in gold. It is melted in a wind-fumaoe, see Fig. 64, with soda and borax, and when it contains base metal, with an addition of a little niter, which serves to toughen it. The melt is poured into an ii^ot^nold and, after cooling, cleaned from adhering slag and shipped to the mint.
OEITBRAL ASKAnOEHEIfT OF A GOLD STAUP-HILL
Fig. 91 indicates clearly the course of the ore through the mill, while Fig. 139, taken to the end of the concentrating tables, gives a general idea of such a mill.
Run-of-mine ore enters the mill at the highest level and is dumped into the storage bin A, thence it is withdrawn through a Blake ore-breaker h, which discharges to the feed-bin C. The feed-bin is filled during the day-shift, and is large enough to hold a twenty-four-hour supply. From the bin the ore is drawn off through a regulated sliding gate by chute to the suspended Challenge ore-feeder, and thus is in constant supply to the stamp battery, where it is crushed with an addition of mercury and of water. The pulp, splashing through the battery Bcreena, flows over amalgamated plates e, where the gold is caught. The tailing from the plates unites in a launder and finally falls into a distributing box that commands four concentrating tables. A distribution is made here and one-fourth the flow is supplied by a launder to each table. The tailing from the tables is wasted. The concentrate is collected and shipped for smelting. The method of driving the machines is indicated in Fig, 88. Above the battery runs an overhead track carrying a trolley and a heavy chain tackle by means of which parts can be removed readily or replaced.
In practice stamps vary from as light as 850 lb. to as high as 2000 lb. as on the Rand, South Africa. Many stamps in American practice are c^ 1000 lb. though there are mills having them of 1250 lb. and 1500 lb. The tendency at present is toward superseding them by rolls or by ball-mills, or supplementing them by tube-mills. In such cases coarser screens are put in at the stamps, their duty is increased, and added work is put upon the' tube-mills, which then grind in closed circuit with Dorr or Akins classifiers (see Figs. 29 and 30).
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The Stamp-Mill With Concentration
COHCEnTRATIOn ra STAHP-MULIKG
Methods. — ^Three general methode of concentration are practiced: (a) removal of high-grade concentrates for shipment to smelters; (b) removal of lo^er-grade concentrates for local treatment by cyanidation, or
Flo. 91.— fitamp-miU foUowed by Amalgamation and Concentration.
roasting foUowed by cyanidation ; (c) removal of concentrates for finer grinding and retmniug to the regular pulp.
Syatema. — ^A very complete concentration system, on a complex ore, is carried out by the Goldfield Consolidated, where approximately 6 per cent by weight and 67 per cent by value of the feed is removed at a cost for concentration of 6 cents per ton of ore concentrated, the concentrates then receiving a very successful local treatment. On the 26-02. diver ore of the San Rafael (Pachuca), IJ per cent by weight and 22 per
cent by value ia taken out as concentrate, to be sliipped to melt«r. On the high^irade Esperanza (EI Oro) sulphide ores, concentration removed 1.2 per cent by weight, with 35 per cent of the gold and 16 per cent of the silver. Tonopah concentration removes approximately 15 per cent of the silver; Stratton's Independence removes approximately 44 per cept of the gold; Liberty Bell, 8 per cent of the gold and 20 per cent of the silver,
Caets. — The cost of concentration, per ton of ore concentrated, ranges from 5 to 15 cents — averaging perhaps 10 cents per ton. The apparent saving in the removal of refractory value and cyanides by concentration, is reflected by decreased cyanide consumption and solution contact, and by lower assay value of final plant residue. This must be carefully balanced agunst the higher cost of realization and mechanical loss in shipment of concentrates.
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Chapter Xiii
The Hydrohetalldrgy Of Gold Ores
HnXIHG ORES ni AQOEOUS SOLUTIONS
At the [M«eent time there are two methods by which gold is dissolved from its ore by chemical solvents. In either process the first step is to obtain the gold in aqueous solution, then to precipitate it from the clear filtrate, and finally to get it in the form of a bar or ingot.
The two processes are:
(1) The cklorinalion or Plaltner process, by which the gold is obtained in solution as a chloride by the action of an aqueous solution of chlorine gas.
(2) The Cyanide or MacArthur-Forrest process, in which the solution of the gold is effected by a weak cyanide solution, the dissolved gold then being present as potassium auro-cyanide. With certain refractory ores, the activity of the solution is greatly increased by the use of bromine or bromo-cyanogen in addition to the potassium cyanide.
Extraction of gold by means of a solvent in aqueous solution is also practiced where gold cannot be completely extracted by amalgamation. This often is the case with pyrite ores; and extraction can be practiced to advantage, not only where amalgamation is imsuitable, but where smelting is expensive.
Gold in ore occurs in particles of various sizes, both as grains readily seen, and in particles of microscopic size. When the particles are visible, or when the ore shows " colors " upon panning, the gold is called coarse, an<f such particles generally can be recovered by amalgamation. Gold often occurs in finely disseminated, microscopic particles, not visible to the eye, and in films on the surface of pyrite crystals. If the ore can be ground so fine as to unlock the crystals, or if it is permeable to solutions, gold can be dissolved in aqueous solvents, such as chlorine or potassium cyanide. Advantage is taken of the solubility of the released gold particles, and leaching or percolation methods, in tanks or vata, are practiced with this in view. The solution soaks through the ore, comes in contact with gold particles, and dissolves them, or by another process, the finely ground ore or slime is a^tated with the solution, and the pulp is filtered and washed in filter-presses. The clear filtrate, in any case, is treated by a
134 The Hydrometallurgy Of Gold Ores
suitable precipitant to obttun the gold in email bulk, and the precipitated gold ia melted and cftst in the form of a bar or ii^ot for sale.
There are three stages in any method of extracting gold by aqueous solvents: (1) The ore is finely ground and when refractory, roasted, to convert the gold into a soluble form, and render it accessible to the solution. (2) The gold ia extracted from the ore by means of a dilute solvent, using a tank with a filter-bottom, or agitating the ore, pulverized to a thin pulp, uwng a filter-press for the separation of the solution. (3) The gold in the solution ia precipitated (a), in chlorioation by hydrogen sulphide or other precipitating agent or (b), in cyanidation by the use of zinc-shaving or sine-dust. The precipitate is collected, dried, and melted into an ingot.
Cyanidation has proved to be a remarkably cheap and efficient method of extraction, but it has limitations, not only in respect te the solubility of the gold, but because of the interference of compoimda that sometimes are present, notably those of copper, that interfere with extraction in various ways. The process has the advantage over the chlorination method in that silver, as well as gold, can be extracted. Under favorable conditions the extraction is high, and modera methods have reduced the cost of treatment to a low figure.
Pyrite ore, exposed to the weather, becomes acid in reaction, and if treated by cyanide, decomposes and destroys the potassium cyanide. To correct this, the ore is first treated by a wash of dilute caustic soda or of acid mixed with caustic lime in sufficient quantity to overcome acidity, or to create " protective alkalinity."
When ore is refractory and requires preliminary roasting, this adds much to the cost of treatment. In chlorination, roasting is always necessary, and in any case it improves the condition of the ore and makes it porous and permeable when leached or filter-pressed.
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Chapter Xiv
This consiBts in attacking the gold in the roasted ore with chlorine to form the soluble gold chloride, and diaaolving out the gold chloride in water.
Orbs Suited To Chlorhtatioh
An ide^ ore for chlorination is one in which the gold is present in a fine state of division, in which bases are absent that would be attacked by chlorine, and silver if present in such a condition as not ta coat the particles d gold with insoluble silver chloride. While the cyanide process is better for the treatment of low-grade ores, many refractoiy high-grade ores have given better results by chlorination.
Ores in which the gangue consists of hydrated iron-oxide are extremely difficult to amalgamate. Not only is the gold finely divided, but the ore is slimy and forms a coating on the amalgamatlng-plates. 8uch ores give satisfactory results by barrel chlorination. Silver is not recovered by chlorination, since it becomes an insoluble silver chloride. If, however, sufficient silver be present to pay the increased cost, salt may be used in roasting and the silver extracted by means oS sodium hyposulphite or better by oyaniding.
Ore containing sulphur, arsenic, and antimony is crushed to 10-to 30-mesh size, and is roasted to expel these elements, to oxidize the bases, to leave the gold in such form as to be attacked by chlorine, and to make the ore porous, accessible to chlorine, and more easily leached.
Chlorinatioa of Cfmcentrate. — ^This is used 06 concentrate from gold-milling, containing much sulphide, and typical of California ore. The coarse gold has been removed, by milling and amalgamation, and the concentrate, generally 1.5 to 2 per cent of the weight of the ore milled, contains gold in fine particles. It is roasted, generally in a long-bedded reverberatory furnace, see Fig. 70, 60 ft. long, and of 3 tons capacity in twenty-four hours. It contains copper, lead, lime, and magnesia, all of which consume chlorine, and form chlorides. To prevent this, it has been customary to add salt, to the extent of 0.75 to 1.S per cent of the charge, at or near the completion of the roast. If roasting has been thorough up to this time, copper is present as CuO, lead as PbSOi, lime as CaO, and B MgO. Were the copper present at the end as CUSO4 it would
react with the salt, forming a chloride of copper. The common salt also reacts upon the gold and forms gold chlorides. Both these chlorides are volatile, and the CuCl, in volatilizing, promotes the detrimental volatilization of the gold.
As long as sulphur is present it protects the gold from attack, but when sulphates have been formed, and are causing the abundant evolution of chlorine by reaction with the salt, the escaping gas carries gold chloride, the gold being unprotected by sulphur at the time of chlorination. Lead sulphate similarly reacts with salt, forming lead chloride, which does not consume chlorine. When much lead is present, however, it may be removed by leaching with hot water before treating with chlorine. Lime and magnema are converted by the salt into chlorides, and in this form consume no chlorine. The process of roasting is therefore conducted as follows:
The ore is thoroughly roasted at a low-red heat. The temperature is a bright red (850° C), to decompose copper sulphate. The salt is added and thoroughly incorporated, and the temperature reduced to prevent volatilization of the gold. The quantity of salt to be added, the time needed for roasting, and the temperature compatible with the minimum loss iA the gold, should be determined experimentally for each kind of ore.
The Goldfield Chlorine He-L Co., Goldpield, Nbv.
At this, the latest development of vat-chlorination, the ore of the 100-ton plant, crushed to 14-mesh. receives an oxidizing roast in a muffle type rabble-roaster, is cooled and moistened, and is then delivered to a storage bin. From the bin it is charged into any one of the seven wooden leaching vats, 22 ft. diameter by 8 ft. deep, by means of a 3-ton grab-bucket operated by a traveling crane. When full, a wooden cover is put upon the vat. and the leaching is done with a strong solution of 8 lb. of chlorine per ton of water. The filtrate from the vat is pumped to a solution-storage tank and is thence drawn to precijatating boxes (resembling the zinc boxes of a cyanide plant), where the gold is precipitated electrolytically. In these boxes are suspended anode plates of grapbitized carbon and cathode plates of lead, the lead having been alloyed with 1 per cent of zinc, the zinc hastening the corrosive action. The resultant slime of the electrolysis consisting of lead, gold, and any silver, while still moist, is mixed with a suitable flux and made into briquettes. These latter are melted in a reverberatory furnace as rich lead bars, are refined m an English cupelling furnace, as described under " Refining Base Bullion," page 498. The pumps and [Mping of the mifl are made of rubber to withstand the corrosive action of the chlorine.
Ma>ri"g the Chlorine Solution. — ^This is produced electrolytically from a
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Barrel Chlorination
brine solution. The chlorine gas, arising from the electrolytic tank, passes to two sheet-iron stacks lined with glazed eewer pipe with cement between the pipe and the outer iron. Inside these is placed a filling of perforated hollow tile balls. Water is supplied to trickle downward through the balls, while the chlorine gas admitted at the bottom is absorbed by the water spread upon the extended surface of the balls.
Barsbl Chlorination
This process was evolved as being better suited to large tonnages of ore than the vat process, and its flows-sheet is shown in Fig. 92.
An example of such an ore is that of Cripple Creek. The ore contains gold telluride, and must be roasted to release the gold from combination
Fio. 92. — Flow.«heet of Barrel Chlorinatioa.
. with tellurium, and to expel all sulphur above 0.1 per cent. The complete process of barrel-chlorination is as follows:
Crushing and Roasting. — The coarsely crushed ore from any of the storage-bins is dravm off as needed to the feed-hopper of the dryer. It is dried and fine-crushed.
Cripple Creek ore is roasted in a mechanical furnace, such as the Edwards, Figs. 74 and 75. The finishing temperature should not be higher than necessary to break up the sulphates formed in roasting.
The cooled ore is raised by aii elevator to stor^^bins, whence it is drawn as needed to the chlorinaticoi barrels.
The ChlDrination BarreL — This is shown in perspective in Fig. 93, and in transverse and longitudinal section in F^. 94.
Within the barrel for a filter a perforated 2-in. plank floor, is used. On the perforated floor rests a lead sheet of 4 lb. per aq. ft. with 0.05-in. holes, I in. between centers. To hold down the filter-sheet, a wooden frame or grating is placed upon it. This is held by blocks h, and heavy
Strips t, securely bolted to the barrel. The wood fnunea beneath last three months; those above, but two or three weeks. This wood-work, if immersed in boiling tar or asphalt until thoroughly impregnated, lasta longer and abBorbs but little solution. The conunon size of barrel is 6 ft. diameter by 12 ft. long, and the capacity is 8.5 tons.
Chafing Oie Roasted Ore.— Into the cylinder is run 800 gal. water, the cbai^ of 8 tons and 8 lb. of liquid chlorine. Chlorine can be obtained in this form in strong steel cylinders or drums. The chai^e-openingB of
Pio. 93.— ChlorinatioD Barrel.
the barrel are now closed, and it is rotated at the rate of 12 R.F.M. for a period of three hours. The chlorine roasts thus:
To see if the saturation with chlorine is complete the stopcock 3 is opened and the isauing gas is tested with ammonia, which produces a white fume with chlorine.
If needed the barrel is stopped and more chlorine is added.
The precipitate collects upon the bottom of the tank, and after several charges have been treated, the united precipitate is drawn off at D and delivered through the man-hole L into the pressure tank z by the hose y. The precipitating tank is then washed clean with the aid of a hoee. The pressure-tank is 4 ft. diameter by 4| ft. high. When charged the cover L is clamped in place, and compressed air, under a pressure of 40 lb. per sq. in.,
Chlorination Barbel
o ° I ■"'
O I /
is admitted throi^h (. At the same time connection is made to the filtei preBB T through the pipe u, and the precipitate collects in the press unde the above pressure. This filter-press is more clearly illustrated in Fig 112. The filtrate from the press passes over a sawdust filter-bed. as t safeguard before it is run to waste. The sawdust is collected occasion^iUy, and burned, to recover the small amount of gold which it may have caught
The precipitate of gold sulphide also contains sulphur, and sulphides o( arsenic, antimony, copper, and silver, forming a " sulphide sake," The press ifl next opened and the precipitate withdrawn.
Compressed air, entering by the pipe w and the valve c, drives the gae through the pipe v and the lead pipe r into the solution in the tank, and precipitates the gold as follows :
The gold is thus thrown down as an auric sulphide in a solution containing both sulphuric and hydrochloric acids. The reaction is rapid, taking about ten minutes.
At first, HaS is oxidized by the chlorine, thiis:
Sulphuric and hydrochloric acids are formed by the reaction, after which auric sulphide is precipitated as in Ek^uation (2).
Filtering.- — After being precipitated, Au2Sa ia allowed to settle two hours. The clear solution then is drawn off at C, 10 in. above the bottom of the tank, through the pipe h into the filter-press. This is done to recover any possible flakes of gold sulphide that failed to settle in the tank x. In three or four hours after precipitation, the tsjik can receive a fresh charge of gold-bearing solution.
After saturation with chlorine, the barrel is revolved for an hour, then stopped in position for filtering with the filtering floor down and level. The outlet pipe k is coimected by a hose to the settling tank and opened; and water is pumped into the barrel above the charge through the valve j. The solution is now drained off, and the water above the charge forced rapidly through by means of compressed air introduced through the valve j. The excess of chlorine is absorbed by the wash-water, and does not enter the building. The operation of filtering is next suspended, connections are broken, valves closed and the barrel revolved a few times to mix the contents again, and to break up channels that may have formed during the leaching. The barrel is then stopped, water run in, compressed air admitted, and the washing resinned. This ia repoat*'d until no gold ia found in the escaping filtrate when tested. The compressed air admitted is under a pressure of 40 lb. per sq, in. The time of filtering and
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Chlokination Plant 141
vashing on an average is 2} hours. The water used is 50 per cent the weight of the ore. All connections are finally broken, valves closed, and manholes opened, and the cylinder is revolved several times to discharge the contents. It is then washed out with a hose to prepare it for another chai^.
ConcentratJon. — ^The washed tailings are, before being discarded, sub-
FlG. 95.— Precipitation Plant f<)r Barrel Chlorination.
jected to a table concentration to recover any particles of unroasted sulphides or tellurides which would contain gold.
Clarifying. — The filtrate is nm through a clarifying press and sent to the stock tanks. From this it is pumped through the opening A, Fig. 95, into the precipitation tank, X, which is 10 ft. diamet«r by 12 ft. high.
Precipitatioii. — We are now ready to precipitate the gold by passing in HaS. To do this, the pipe v is connected to the lead-lined generator G which contains lumps of iron sulphide resting on a perforated lead false-bottom. Dilute sulphuiic acid admitted below the false-bottom comes in
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cfmtact with the iron sulphide and abundantly generates HaS according to
thee
This method of treatment, sucoesBfully conducted at the Colorado ^ringB plant, was abandoned in favor of the cyanide method in 1912.
When the ore is refractory and requires preliminary roasting, this cost adds much to that of treatment. In chlorinatioD, roasting is cfmunon, and in any case it improves the condition of the ore, and makes it porous and so permeable when leached or fiherftreseed.
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Chapter Xv
This is a hydrometallurgica] method of treatment, that a, tbe gold is recovered in a weak water solution of cyanide.
GRAVTTT CONCENTRATIOH PRIOR TO C^AniDIHG
This precedea cyanide treatment, on account of being more conveniently applied at this point, but in many cases it would be advantageous if it were feasible to have it follow cyanide treatment. When it precedes cyanide treatment, a considerable proportion of the readily soluble gold and silver may be removed in the concentrate, when it might be more advantageous to recover them as bullion. In the case of low freight rates and favorable smelter contracts this may be an advantage rather than otherwise ; but for the majority of plants the more gold and silver turned out as bullion, the better. Therefore, the ideal practice in most cases would be to recover all the gold and eilver possible as bullion by the cyanide process and then concentrate the tailing to recover as high a percentage of the remaining gold and silver which had escaped dissolution as feasible. Obviously there are many cases where the percentage of extraction of gold and silver from the residue leaves no mat^ for concentration.
At times concentration can be introduced into tbe cyanide flow sheet to advantage for the purpose of removing the minerals which are difficult, if not impossible, to treat by the regular mUUng scheme, so that they may receive the special treatment necessary without having to incur the expense of subjecting the whole tonnt^e to the special treatment made necessary by a comparatively small proportion of refractory minerals. On the other band, the introduction of concentration adds to the first cost of the plant and the complexity of its operation, so that these disadvantages must be carefully weighed against the advantages which are likely to accrue.
OUTUNB OP THE PROCESS OF CTANTDmo
The ore is crushed to such fineness that its contained gold is left open to tbe actioD of a weak solution of potassium or sodium-cyanide. The gold, brouf^t into solution, is then precipitated from the clear filtrate and this preci[state melted into form of a bar or ingot. The gold in tbe
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cyanide sohitioo is present as potassiiun or sodium auro-cyanide. With certain refractory ores the solvent power of the solution is greatly increased by the addition of bromine or bromo-cyanogen.
Extraction of gold by cyaniding ia successful, where gold cannot be completely removed by amalgamation. This is the case with pyrite ores, and Euch extraction can be practiced to advantage, not only where amalgamation is unsuitable, but where smelting is too expensive.
Practically all gold or silver ores cwi be treated by the cyanide process or by the cyanide process in combination with some other preliminary or accessory treatment, ^th the exception perhaps of certain ores containing a considerable amount of copper, lead, etc. In such cases, the desirability of recovering the base metab leads to smelting the ore upon either the lead or copper basis, and the necessity for hydrcmetallur^cal treatment for the recovery of the gold and silver disappears since the precious metals are recovered by the smelting operation in connection with the base metala.
Cyantdation has proved a remarkably cheap and efficient method of extraction, but it has limitations, not only in respect to the solubility of the gold, but because of the action of certain compounds, notably those of copper, that are sometimes present and interfere with extraction in various ways. The process has the advantage over the chlormation method in that silver, as well as gold, can be extracted. Under favorable conditions the extraction is high, and modem methods have reduced the cost of treatment to a low figure.
The Use of. Hot Solutions in Cyaniding for Better Extraction of Silver and Gold. — At the Belmont mill, Tonopah, the temperature at the stamps is 60° to 70° F., and by the use of exhaust steam at the Fachuca agitators, the pulp is raised to 90° to 100° F., resulting, as reported, in increased extraction over using cold solutions of 2 per cent. At the Montana-Tonopah mill crushing is done at 50° to 60° F,, and by live steam in the agitator, the temperature is brought up to 1 10° F, It is found at this mill that when heating is not done extraction falls off, also that heat aids settling. At the MacNamara mill, Tonopah, the pulp was heated to 115° to 120° F., using live steam in the agitators, whereby, as compared with cold solutions the extraction increased and the time of agitation was lessened. Even as compared with a temperature of 80° an increase to 120° improved extraction by 1 .5 per cent to 2 per cent. The cost of this heating may be reckoned at 18 to 30 cente per ton as against a savbig at 2 per cent or 60 cents per ton on dollar silver. On gold ore savings are not so secured. At Kalgoorlie, Western Australia, where ores are roasted, the temperature, after mixing with solution, is as high as 200° F., but they prefer to cool the ore before mixing.
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Cyaniding Ores
ORES SUITED TO CYAinDATIOW
Dealing with gold-bearing ores, the following claas^ are amenable to treatment:
1. Talcose or Clayey Ores. — When crushed, these produce a high percentage of slime, which is too fine for leaching, and gives trouble in any type of filter. Generally, the capacity of a plant treating these ores is not high.
2. f^ve-ffiilting SilJcious Ores. — These constitute the bulk of the ores treated by the cyanide process throughout the world. The gold may be fine or coarse, in the latter case being removed by amalgamation before subsequent cyanide treatment. It is not economical to dissolve coarse gold in cyanide.
3. Pyrite Ores. — Gold in this class is in most cases mechanically mixed with the iron pyrite, which when crushed, liberates the gold for the solution in cyanide. It has been found in several mining districts that it is possible to treat the pyrite mixed with the ore, no concentration being necessary; but in the majority of cases, it is found better to concentrate, and treat this product separately.
4. Telluride Ores.— These occur at Cripple Creek, Goldfield, Kalgoorlie, and other places, and while not complex, have given considerable trouble in treatment. Ordinary cyanide solutions are not effective on tellurium compounds, and the ore should be either concentrated and treated by ordinary cyanide or bromo-cyanide, or all the ore roasted, followed by the usual cyanide methods.
5. Antimony Ores. — Antimony is a troublesome mineral to de^ with in gold extraction. It occurs in gold ores in Rhodesia to some extent, and in New South Wales, Roasting seems to be effective, while caustic soda solutions have helped cyanidation. Tailing containing antimony has been successfully treated in Australia with no special process.
6. Graphite Ores. — At Ashanti, West Africa, Kalgooriie and Gympie, Australia, graphitic slate or schist is mixed with the ore, and while not containing a high percentage of gold it causes a premature precipitation of gold from cyanide solutions.
7. Copper Ores. — Many gold ores contain a low percentage of copper, and with care may be treated with a fair recovery. Copper gradxially changes cyanide solutions, and then is precipitated on the zinc shaving, preventing a proper precipitation of gold. Copper in the resultant bullion with care may be refined.
8. Arsenical Ores.— A mineralized ore often contains a little arsenic, but pure mispickel, or arsenical pyrite, requires skill in treatment. The ore may be roasted after fine crushing, or concentrated, and the product roasted alone. Noted cases of mines producing this ore are at I
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in Victoria; the LanceSeld and Transvaal mines, in Western Australia; the Deloro and Hedley in Canada. From the first mentioned group, the ftyiite is roasted and treated by cyanide or cblorination; at the Lanoefield both wet and dry processes have been tried on lai^ tonnages; the Ttbobvaal mine ore is very refractory; while at the Deloro, bromo-cyanide was Tised for several years. At the best, it may be said that such a gold-bearing ore ia difficult to treat.
Chbhbtrt Of The Cyamdb Process For Oold Orbs
When a solution containing from 0.1 to 0.5 per cent cyanide is broi^t into contact with crushed ore containing very fine gold, this metal is eaaly dissolved. According to Eisner, the equation is as follows:
The gold is dissolved by the action of potassiiun or sodium cyanide in the presence of oxygen and water, forming an auric-potassic cyanide and caustic potash. Oxygen is needed to fulfill the requirements of the reaction, and consequently ore, or solution acting on ore, must be aerated in some manner. When oxygen of the dissolved air is consumed, action ceases, but resumes with a fresh supply of air. Oxidising agents such as potasdum chlorat« and permanganate, and the peroxides of lead, manganese, sodium, and barium may be used to furnish oxygen in place of air, but have been found too expensive for practical use.
The inadequacy of this equation as a guide to consumption of cyanide is at once apparent in the treatment oS complex ores, with their various constituents. The equation calls for 1.51 units of Au or 0.83 unit of Ag per unit of KCN. With many silver ores this relation is closely approximated, but with gold ores a consumption of 20 of KCN to 1 of Au is considered satisfactory, and 40 to 1 is more common. The mixed potassiimi cyanide salt (98 to 99 per cent KCN) has been quite generally supplanted by sodium cyanide (120 to 129 per cent, in terms of KCN). Subsequent references to cyanide will be in terms of 100 per cent KCN, although sodium cyanide is used.
When an ore containing pyrite is exposed to the weather, air and moisture slowly act on the mineral, with the following reaction :
(2) 3FeS2-|-2HaO-|-220-FeS04-|-Fe2(S04)3+2H3SO*.
Ferrous and ferric sulphates and sulphuric acid are thus formed. The first two named would tend to precipitate gold. Ferric sulphate is acid in its reaction, and with sulphuric acid, if not neutralized, it would decompose and cause a serious loss of cyanide. Such compounds are called " cyaniddes," Ore, therefore, which contains pyrite, and has been exposed to the
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.Reactions In Cyaniding 147
weather needs caustic soda or lime to neutralize the addity, and an excess to provide for any acidity resulting from further decompomtion. This excess is termed the " protective alkalinity." To remove the soluble ferrous sulphate and sulphuric acid, water-wash before treatment should be sufficient, but would require some time, so a certain quantity of lime is added. Lime is cheaper than and preferable to caustic soda, the latter making UDdearable compounds, often noticed later in treatment. In some districts, the question of freight will decide which should be used. The action of lime is as follows:
The result is the formation of a harmless iron hydroxide and calcium sulphate.
The reaction that takes place, when a gold-bearing solution comes in contact with sinc-ahaving in the precipitating boxes, or when nnc-dust is mixed with it is :
(6) KAu{CN)3+2K(CN)a-|-Zn+H20-KaZn(CN)4.-|-Au^KOH+H
in which one part of zinc is computed to precipitate three parts of gold. The gold forms a brown or black precipitate and the rinc potassic cyanide reoiains in solution.
Cyanide is also used up by direct combination with the zinc aa follows:
In both the foregoing reactions bydrc^en escapes in bubbles. When alimiintmi is used we have:
in which one part of aluminum precipitates 7.2 parts of gold.
The barren solution, from which the gold has been extracted, is used again in the mill, and accumulates impurities from ore that is being treated, and from the zinc with which it was in contact in the zinc-boxes. As a result, it gradually becomes less efficient than fresh solution. It has been found that, on adding lime to a cyanide solution, its solvent power upon a clean ore is increased, but not on a sulphide ore. Such a solution, if treated with sodium-sulphide to the point of exact neutrality, and with a small
excess of lead acetate, and given time to permit the resultant sulphide to precipitate, is improved in dissolving power as follows:
The cyanide b here regenerated, while the zinc sulphide separates. This is a means of overcoming the accumulation of zmc in solution, which is one of the drawbacks to the use of zinc for precipitation, compared with electrical deposition. Chemicals, however, are not indispensable for disposbg of the zinc. In Eisner's equation, caustic potash is set free. This reacts upon the sulphides in an ore, forming soluble sulphides, which in turn react like the sodium sulphide in the reaction above, precipitating zinc sulphide from the ore.
The following minerals and chemical compounds destroy or combine with cyanide, and render it incapable of dissolving gold: Copper in the form of sulphate, carbonate, copper glance, erubescite, or copper pyrit«, (The sulph-antimonites of copper are without action.) Manganese as " wad " (impure hydrous oxide), but not the carbonate or oxide; zinc as smithsonite, but not blende or zinc sihcate.
Graphite, which is found in certain ores, also carbon remaining in burned lime, both interfere with extraction by causing a premature precipitation of gold. Also leaves, roots, and other organic matter act in the same way.
It has been found that gold thus prematurely precipitated is soluble in a solution of sodium-sulphide, this being added when leaching the sands after practically all the cyanide has been displaced by a water-wash. Precipitation of the gold is effected by passing the NaaS through boxes filled with copper shaving, but little copper going into aolutioti.
To increase the activity of zinc -ihavbig in precipitating gold, it may be dipped in 10 per cent lead acetate solution, or a drip of the latter may be fed in at the head of the zinc-boxes. This forms a zinc-lead couple, which reacts electrically on the gold solution. Both potassium and sodium cyanide, 98 and 128 per cent pure, are used with varied results, the latter being rather more favored as a dissolving agent.
The cyanogen contents of potassiimi and sodium cyanide respectively are about 38 and 51 per cent, so for the same weight of wait there is the extra percentage of cyanogen, a consideration where fre^ht is costly. Commercial cyanide contains small quantities of alkaline sulphides, whose presence diminishes the solvent power of the cyanide. It pays, therefore, to buy the salt on a guaranteed analysis.
The Concentration of the Cyanide Solution is a vital point. In general, the stronger the solution, within certain fairly well-defined limits, the more rapid the dissolution, and, furthcnnorc, the loss the interference of a given percentage of impurity which might i>e present in solution. On the other hand, a greater proportion of impurity may be dissolved by the stronger
The Bromo-Cyanide Process 149
solution. In the case of a plant which is operating at forced capacity a stronger solution may be used to advantf^ in order that the maximum extraction may be attained under this condition of operation. Stronger Bohitions may result in increased cyanide consumption, although the magnitude of this loss in a properly operated plant is not so great as has been supposed.
However, in a plant where there is considerable mechanical loss of solution the additional cyanide loss would prove an important factor. For tiiis reason a strong solution is not generally favored when continuous decantatjon is used. The tendency in some cyanide plants has been to use a lower concentration in cyanide than that capable of giving the highest economical result. This probably is through the fear of excessive cyanide loss.
The Bromo-cyanide Process. — Tellurides of gold and silver are practically insoluble in plain cyanide, but are soluble in bromo-cyanide solutions. The latter process was first used on a large scale in 1899 at the Hannans Star mill, Kalgoorlie, which was mainly a customs plant receiving many shi[»nents of rich telluride ores. The process involved is commonly known as the Diehl.
The bromo-cyanide solution is made according to the following equation :
(10) 2KBr+KBr03-|-3KCN-|-3H2804 = 3Br€N-|-3K2SO»+3H20 Its action in the treatment vat is supposed to be as follows :
(11) BrCN-|-3KCN-t-2Au=2KAuCN3-|-KBr.
The first two quantities in Equation (10) are contained in the mixed salts having about 40 to 44 per cent Br as KBr, and 20 to 22 per cent Br as KBrOs; the proportion of Br as bromide being about twice that of Br as bromate. A 30-lb. charge is usually made up, SO lb. of 63 per cent H3SO4, 20 lb. of KCN of 93 per cent and 36.S lb. of mixed salts as above given.
The solution is made in a closed wooden vessel, holding about 200 gal., stirred by rotating arms. In making up a charge, a portion of the water and all the H2SO4 are first mixed, and allowed to cool to normal temperature. The KCN, which is dissolved in a separate vessel in sufficient water to fill the mixing vessel, is then run in, and at the same time the proper weight of " mixed salts " is gradually added. The whole is then agitated for six hours before being used, and in a closed vessel it will retain its strength for some days. The cost of a 30-lb. charge of BrON is about ^1.60.
Ebtperiments have shown the following points necessary for good work :
1. The daily ore sample should be taken in the morning, and assayed
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as soon as possible, so that the value of the ore pasBing to the vata in the previous twenty-four hours may be determined.
2. The pulp should have a long KCN treatment.
3. A vat should be kept under KCN treatment until the vahie of the KCN residue is known.
4. The alkalinity of the vat should then be determined and corrected to 0.01 per cent by HjSO* before adding BrCN.
5. The quantity of BrCN added should then be detennined from the value of KCN remdue, and the tonnage of the vat.
6. The lime added to the ore during crushing should be varied according to the alkalinity-test after KCN treatment, so that the plant-eolution tests about 0.02 per cent.
7. Lime water should be made and added to the vats or to the solution fran the presses, instead of adding lime to the vats.
8. Metallic iron should be kept out of the pulp as far as posable, as it is both a cyanicide and a bromo-cyanicide.
In operation, a vat, when full, was given ita charge of KCN, and three hours afterward a " dip " KCN residue was taken, and the charge of BrCN solution added. After a total a^tation of twenty hours a qufuitity of lime was added, and the vat-chai^ " pressed." The quantity of BrCN added was varied according to the residue of preceding vats, and the value of the ore being treated as shown by the daily ore-sample. Each charge of bromo-cyanide was totally destroyed. In places where fuel and furnace supplies were expensive, this chemical process would show a decided advantage. The process requires more metallurgical skill and constant attention to the progress of each vat under treatment; but, if this is available, the results are highly satisfactory, and the process has deSnite claim to be a cheap and efficient method of treatjnent for such ores as the Kalgoorlie sulpho-tellurides.
THE STANDARD STSTEHS OP CTAHIDATIOff
These may be divided into three as follows:
1. Sand leaching, where the whole ore pulp after grinding is classified into two products, sand and slime, the sand being sent to percolation tanks for leaching; the slime or fine product being filtered or decanted to yield a pregnant solution and a solid residue.
2. Filter slime treatment, where the whole pulp is ground fine and filtered to yield a pregnant solution and a solid residue.
3. Slime Agitatioi, where the whole pulp is ground fine and agitated for a considerable period, then by countercurrent decantation yields a pregnant solution and a barren or nearly barren residue that is thrown away.
Practically the " sand " is the portion of the pulp treated by pertxdar
Sand Leaching 151
Hod or leaching; while the balance is considered to be slime and is treated by suction or pressure filtration.
With the important exoeptitns of the Rand and the Homestake the earher practice of sand leaching, that is of comparatively coarse grinding and treating the sand and slime separately has largely given way to the present practice of (2) filter slime treatment or (3) slime Station (both " all-sliming " treatment) the pulp being finely ground and treated as a single product.
Referring back to the so-called sand leaching, the pulp, rather coarsely ground, is classified in cone classifiers such as the Caldecott giving a sand for percolation and a slime such that 90 per cent of it will pass SOO-mesh (0.0029 in.).
In aU-sUming plants the pulp varies from 60 to 90 per cent throi^h 200-me8h, depending on the economical limit of fine grinding, and subsequent treatment. In sand and sUme plants, crushing b almost universally done in water, while with all-sliming, crushing in cyanide solution is almost univental.
Separation irf Sand fn«n Slime. — This is an interesting problem, and is performed either to furnish from the pulp a sand for fine grinding or again to make two products, sand and slime, the sand to be leached into vats, the slime to be separately treated. The machines used in separating the sand from the slime and their principles of operation are fully described under head of " Classifiers."
Leaching the Sands. — Cyanidation was first apphed to the recovery of gold in accumulated tailii^s from stamp-mills. In South Africa this material still containing $3.50 per ton, was impounded or retained behind dams until cyaniding could be undertaken. The tailing was shoveled into cars and hauled to large leaching vats. Here the ore was leached with weak cyanide solution, the gold precipitated from the filtered solution by passing it throi^h boxes containing zinc shavings. The precipitate was treated, melted, and obtained in form of gold ingots or bars. This practice, as long as the impounded tailings lasted, was quite simple, but it has been abandoned with the exhaustion of those accumulations.
Description of Vats. — Leaching vats are constructed with filter bottoms and may be made of wood or steel. In warm countries, like Australia, South Africa and Mexico the steel vat is preferred ; but in cold countries, where it is necessary to house the plant, the wooden vat gives satisfaction. The latter is cheaper in first cost and easier to set up, though the wood absorbs gold solution. The steel vat, on the other hand is less liable to leak, but should be painted. A wooden vat, also, when requiring it, should be painted. In Western America wooden vats predominate.
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Fig. 96 is a perspective view of a wooden vat with tlie filter-cloth omitted. This shows the false-bottom of slats and the hinged bottomdischai^ opening through which the exhausted tailing is shoveled or sluiced
Pia. 96.— View of Wooden Leaching Vat
out. A wooden ring, 2 in. high and 2i in. thick, is nailed to the bottom d the \at, leaving a space of J in. between it and the side. Parallel strips, I in. high, are nailed a foot apart upon the bottom, and across these 1-by
Fio. 97.— PlftD of Steel Leaching Vat.
4-in. strips or slats are laid with 1 in. space between. Upon this false-bottom cocoa-matting is spread, and over it 8-oz. canvas filter cloth cut 12 in. larger in diameter than the vat. The edges of the cloth are held
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Double Treatment 153
down by a rope laid upon the canvas and driven into the }-in. space between the stav^ and the wooden ring.
Figs. 97 and 98 represent in plan and in elevation respectively, the construction of a steel vat having a perforated board bottom. A ring of flat iron ^ by 2) in. is riveted to the side of the vat, with space* thimbles to hold it ^ in. from the side. The cleats that sustain the false bottom are 2 in. high by Ij in. wide. The 1-in. bottom-boards are bored with ^-in. holes and screwed to the cleats. As in the case of the wooden vats, ,^y
the thick stiff cocoa-matting is „ „ .„,,„,
■ . J 41. * 1 I 4. J P'°- 98.— Sectton of Steel Leaching Tank,
laid upon the false-bottom, and
covered with a filter-cloth of S- to 10-oz. canvas. The edges are calked with
J-in. rope into the }-in. space, as shown aXgia the sectional view, Fig. 98.
Leaching vats vary in size from 16 to 50 ft. in diameter and 4 to 9 ft. in deptii. The shallow ones are for the more finely ground sand.
Double Treatment. — Fig. 99 shows the two steel vats or intakes used in this system as practiced in South Africa. It consists of an upper or settling vat 40 ft. diameter by 7\ ft. higli to which the sands are conveyed to be spread out and to receive a preliminary teaching, and of a lower or leaching vat 12 in. deeper. Both vats are carried on a steel structure for access both above and below. When the leaching is sufciently completed in the settling vat, then seven bottom doors or valves are opened and the material is shoveled into the leaching vat beneath, 40 ft. by 8i ft. h^h. When thus ^mn handled, the tailings become more bulky, and more open and even for leaching. Here leaching and washing is completed. The exhausted tailings are then withdrawn at the dischai^ openings similar to those of the upper tank, and trammed away to the waste diunp.
Leaching the Sands. — While transferring the sand from collectors to leaching-vats, lead acetate, previously dissolved in water (§ lb. per ton) is added and slacked lime (4 lb. per ton) is thrown into the collecting vat, thus thoroughly mixing the lime with the sand. After transferring, a little shoveling is done to level the sand. The first leaching solution, amounting to 30 tons, is brought up to 0.25 per cent strength by the addition of a sufficient amount of potassium cyanide solution of known strength to the vat under treatment. This amount of strong solution is allowed to drain slowly through the partly opened drain-valves and is followed by repeated washings of weak solution from 0.15 to 0.20 per cent, after which
Cyaniding Of Gold Ores
the charge is drained (or transfer. This first treatment occuines five days, including time of the latter.
Outside Elevation
Fio. W.— Vata for Double Treatment.
sizmo TBST on sabd REsrottB
Remaining on. Remaining on. Rem&iniDKon. Remaining on. Remaining on. Remaining on. Remaining on.
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Sand Leachinq 155
The second treatment aver&ges five days and con»ste of repeated washings of strong and weak solutione, that are drained oS, and the sand teanHfened to another vat for the final treatment, which consists of aa many washes of wash solution as there is time to apply, followed by two or three of water to displace all the solution. Then the vat is finally drained by vacuum for discharging from the plant. Each chaige of solution is allowed to disappear below the surface <rf the sand before the succeeding one is applied. Sand undergoes treatment for twelve to fifteen days.
All leachingB from the sand vats, as well as the plant solutions, are sampled, assayed, and titrated for cyanide and alkalinity daily. Attenuated leaching solutions are sent direct to weak sumps. Centrifugal pumps, when not pumping to treatment-vats, are in service circulating solution in sumps through cones, Fig. 58, for the pmpoee of aSrating.
All potassium cyanide used in the treatment of .sand, is dissolved in a small vat from which a 2-in. pipe-line is connected to the suction of a 4-in. centrifugal pump used to pump solutions on sand. By means of a table and float arranged on the vat, the desired strength of solution can be obtained by opening the 2-Jn. line and allowing the requisite amount of standard solution to be drawn through the pump with the weak solution from the weak sumps.
The Rand is the greatest present-day ex;>onent of sand leaching, comparatively few plants in other parts of the world retaining separate sand and slime treatment. The important improvements made in fine grinding and the handling of slime created a marked tendency toward allsliming and the abandonment of sand leaching, this trend, in a few instances resulting in inadequate consideration of the question of highest commer-ci^ recovery versus theoretical extraction. A few all-sliming plants, after careful research and consideration, have found it advisable to revert to the former sand and slime practice. But even in the strongholds of sand leaching, the present day trend is toward finer grinding and a gradual reduction in the proportion of total mill pulp treated by percolation.
The removal of the residue from the leaching tanks, which now rimge up to 70 ft. diameter by 12 ft. deep, is accomplished by sluicing, shoveling, or mechuiical excavatora, the method depending on local factors such as cost and abundance of water, and labor and storage requirements. A recent trend in the method of sand-residue disposal has been to sluice the sand into dewaterers, or classifiers, from which it gravitates under ground, for stope fiUing. The destruction of the cyanide present in the residues is accomplished, where necessary, by the addition of some 0.08 lb. ' of KMn04 per ton of sand. Smce its introduction on the Rand, sand fill- [ iag of stopes has been very advantageously adopted in other countries
Rand Metallurgical Costs. — In this practice, the following comparison of the coet per ton for transferring dnuned sand from tank to tank is: (a)
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by truck haulage, 3.96 cents; (6) by shoveling and conveyor belt, 3.36 cents; (c) by shuttle belts and main conveyor belts, 2.44 cents.
In offering sand-leaching costnJata, it may be well to cite the working of a few representative plants. Again, it must be noted that no comparisons are here intended, local conditions being at too great a variance. Fifteen Rand plants show that an average of 57.6 per cent of the mill pulp is leached as sand by double treatment, the proportion ranging from 70 per cent sand in the older plants, to 40 per cent sand in recent pltuits. The economical limit of fine grinding of the sand approximates 70 per cent through 100-mesh (0.0058-in. aperture). With intermittent collecting, revolving diatributora are preferred to hose-filling. The transfer of the drained sand to the treatment tanks is by truck-haulage, or by shoveling, in cases where the collectors or settling tanks are set above the treatment tanks. With continuous sand-filter collecting, the collected sand is transferred by conveyors or pumped to revolving distributors. The residue is discharged by belts, trucks, buckets, or sluicing into cone-dewaterers for sand filling of stopes. The total ratio of solution to sand for dissolving and washing approximates 1.5 to 1. About four days are allowed for collecting and draining and seven to eight days for treatment. Approximately one ton of solution is precipitated per ton of sand. The strong solution is about 2 lb. KCN, while the weak is about 0.5 lb. KCN,
Figures from ten Rand plants for 1912, with daily sand tonn^e running as high as 3500 tons, show the following averages:
Total coet of sand treatment per ton of aand (inchidmg cUaaification, collecting, treating, precipiUting, refining and residue disposal), ranging
from 37 to ih cents, averftge 39.8 cents
Average extraction on $3 17 sand heads 79 .8 per cent
Average extraction in recent Hand plants 89.5 per cent
Costs of Rand Plants. — The capital expenditures for various capacities of plants, per ton of aand treated per twenty-four hours, are as follows; 350 tons of sand per day, $264; 700 tons, $254; 1400 tons, S244; 2800 tons, $233.
Estimated capital costs of sand plants using continuous sand-filter collecting, yielding 55 per cent aand, are given as follows: 275 tons of sand per day, $230 per ton of daily aand capacity; 550 tons of sand, $212; 1100 tons of sand, $195; 2200 tons of aand, $177.
Cost of Homestake Plant. — ^The actual combmcd cost of the two Home-stake sand plants, of a total capacity of 2500 tons, amounted to $255 per ton of daily sand capacity This figure includes power, heating, precipitating, and clean-up plants.
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Slime Treatment
Decantation. — Aa largely practiced on the Rand, the slime is settled in slime collectors, large eone-bottom tanks provided with outlets for the escape of the water, and in adjijstable decanter, all the slimes settling to the bottom of the cone. The settled slime, containing approximately 50 per cent of moisture, is then sluiced out by aid of a weak cyanide sohition and is pumped to the " first settlement tank." Here the charge is kept in agitation by means of a circulating pump which withdraws from the bottom of the tank to discharge at the top. The charge when sufficiently agitated b allowed to settle and the gold-bearing solution decanted, then precipitated at the zinc boxes. The settled charge of slime in this tank is transferred to the second settling tank with addition of weak cyanide solution, and after settling and decanting of the supernatant solution, the thickened slime is discharged, using for that purpose sufficient fresh water. Settlement is aided by the judicious use of lime. This method, while simple and inexpensive, has the drawback that there is an inevitable loss through the imperfect washing, though for low-grade ores such loss is inconsiderable .
Aa-sliming. — Except in the case of certain low-grade mines, such as the Rand and Homestake, the earlier practice of comparatively coarse grinding, then treating the sand and slime separately has largely given way to the practice of finer grinding and treating the entire pulp. " All sliming " is the term applied to this. Where a separate sand slime treatment has prevailed the two products have been separated by classifier, the slime of such fineness that 90 per cent of it will pass a 200-mesh screen. In the allsliming plants the pulp varies from 60 per cent to 90 per cent through 200-mesh, depending upon the economic limit of fine grinding and subsequent treatment. In the sand-slime plant the grinding is done in water, while where all-sliming prevails, thb is done in cyanide solution.
This is done for the purpose of bringing the slimed ore intimately in contact with the cyanide solution and with air according to Eisner's reaction.
The matter of proper dilution at which pulp should be agitated is one demanding careful consideration for each individual plant. Experimental work points to high dilutions, but practical considerations generally necessitate as thick a pulp as can be regularly drawn from the thickeners. A dilution of 1.5 to 1 is, perhaps, an average. Inasmuch as the cost of agitating a ton of slime for twenty-four hours should not exceed 3 cents, it is evident that the agitation period, to be commensurate with this loir cost,
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15S
Cyaniding Of Gold Ores
should be well advanced toward the point where further dissolutirai ceases. In determining the proper length of agitation, such items as capital charge on the t^tators and accessories, consumption of cyanide, lime, and other chemicals, posrable premature precipitation of dissolved metals in prolonged agitation, and the mechanical cost of agitating, should be carefully balanced gainst the net returns from the metals dissolved. Then, too, the possibUity of decreasing the fineness of grinding by prolon^g the agitation should be examined.
The modem trend is in favor of continuous a^tatJon, as compared with the former intermittent system. Change of solution during fetation is found advisable with certain ores; althou(^ with present-day f^tators, capable of giving ample aeration, the entire agitation step ia frequently carried out in one stage-
The types ai agitators may be divided into (1) pneumatic, (2) mechanical, (3) combined mechanical and pneumatic, (4) pump transfer system.
(I) PNEUMATIC AGITATORS These include the Parral and the Fachuca (or Brown). The latter is an efficient machine for concentrates, but in the majority of cases the cost of pumping the pulp into these tall tanks is one serious thing against them. It is also costly to erect.
The Pachuca or Brown tank, shown in Fig. 100, consists of a steel cylinder ordinarily 30 ft. high and 10 ft. in diameter, terminating below in a cone having an angle of 60° at the vertex. Vats 60 ft. high and 15 ft. in diameter are being constructed in some mills. In the lower part of the cone there is a 6-in. pipe with a gate valve for discharging the contents of the vat after treatment. The apparatus in the interior of the vat consists of: (1) A central 10-in. tube called the elevator. (2) A pipe for compressed air passes through the center of the elevator and rests on the bottom of the cone. This pipe is closed at the bottom end, but has a number of small holes in it at the level of Uie bottom of the elevator tube. These holes are covered by a section of rubber hose slipped Fio. 100 —Pachuca Tank. ^^' **^^ P'P^ a"<^ having its lower end tied to the pipe with wire just below the holes in the pipe, so that the hose forms a sort of collar valve through which the aJr
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The Pachuca Tank 159
may escape into the vat, while the pulp is prevented from entering the pipe. When the piessiii« of air in the pipe is greater than that due to the cohinin of slime in the vat the air escapes through the collar valve and passes up through the end, but has a number of small holes in it at the level of the bottom elevator tube, carrying the slime with it. (3) Another air pipe with its collar valve is placed outside of the elevator to keep the slime in circulation during the filling and emptying of the vat. (4) An adjustable agitating device consisting of an annular pipe having several small pipes with collar valves so arranged that compressed air, water, or solution may be forced through them in order to wash off any sand or slime which may have deposited on the sides of the cone after the agitation has been stopped.
The method of operation to cause the agitation or circulation of the slime in these vats is vefy simple, being as follows : When the vat is filled with pulp and solution, the valve is opened, admitting compressed air into the bottom of the elevator where it mixes with the pulp in the elevator, and as this mixture is lighter than the pulp in the vat, it rises to the top of the elevator and overflows into the vat, while another portion of the slime in the vat enters the elevator at the bottom which in turn is raised to the top and overflows, thus obtaining a perfect and continuous circulation of the pulp as long as the compressed air is allowed to enter. Upon starting the circulation the air pressure must be greater than that of the colupm of alive, but as soon as the circulation is well established lees pressure is required, it having been found in practice that while 50 lb. pressure is required to start the circulation, as soon as the sand and slime which have settled in the bottom of the cone have been cleared out by the scouring of the circulating pulp, the circulation may be maintained by a pressure of but 25 lb. per square inch. The quantity of air required in any particular case depends on the proportion of SMid to slime, the fineness of the pulp, and the viscosity of the slime. Ordinarily, in the Pachuca plants, 100 cu. ft. per minute is used to maintain a vat containing 100 tons of slime in active circulation and to prevent the settlement of the sand on the sides of the cone bottom, but in the Goldfield Consolidated plant, Nevada, and at the Komata Reefs, New Zealand, only 30 cu. ft. per minute is used.
Mechanical Agitators. — For the agitation of heavy sulphides we have the old mechanical agitator with plowshoes revolving at 16 to 18 R.P.M. Much power ia required to operate it.
Top-drive paddle-arm stirrers or agitators were used from the start of the slime-agitation process, and with proper design, still hold a fair place at this day. The bottom-drive agitator, as used in Kalgoorlie, Mexico,
and the Rand, elinmiatee the submerged step beaimg of the top drive, allows of a simpler overhead construction, and due to the increased rigidity of the drive, operates with very moderate power. Assuming a tank 30 ft. diameter by 12 ft. deep as an approximate standard, a satisfactory agitating speed for the bottom drive type may be given as 8 R.F.M. and the power as 6 H.P. In addition, the power for furnishing 20 cu. ft. of free air per minute at 10 lb. pressure, for requisite aeration, amounts to 1 H.P. This type of agitator works quite satisfactorily, although it is troublesome to start up after a protracted shuir-down, tends to bank up near the center and gives en masse agitation. Inclined baffle-boards bolted to the side of the tank will materially assist agitation.
The average erected cost of a 30Xl2-ft. bottom-drive, steel-tank, mechanical agitator may be given as S2395. Such a tank will have an available depth of 11 ft. 6 in., and, with pulp at adilution of 1.5, will hold 134 tons of dry slime, making the total erected cost $17.87 per ton of dry slime capacity.
Agitation of the SUme. — By fine grinding even microscopic particles of gold have been set free from the gangue or waste. These particles can be dissolved provided they can be brought intimately in contact with the cyanide solution in presence of air. This is done by agitation, preferably using an air jet for so doing. In this way the solution is well circulated and is brought in contact with each particle of the slimed material as ordinary stirring or mixing will not do. Even then the solution of the gold particles takes hours of time, and agitation is continued until assay shows that further dissolution has ceased.
(S) Combined Hgchanical And Pneumatic Agftators
These include the Dorr, the Trent, and the Hendryx machines. The Dorr agitator appears to be the cheapest in first cost, requires less power, and will operate with the least amount of trouble. It is well suited for a concentrate, ground to pass a 200-niesh screen. The Trent and Hendryx machines are not well suited for heavy sulphide material.
Dorr Agitator. This comparatively new type of agitator, combining mechanical and pneumatic agitation, has met with merited success. It is usually operated at 3 R.P.M., the speed depending, of course, on the chai^ acter of the pulp. Under average conditions, a 30X12-ft. Dorr agitator will require about 1.5 H.P. for moving the arms, at 3 R.P.M., and 2.5 H.P, for furnishing 30 cu. ft. of free air per minute at 20 lb. pressure, a total of 4 H.P. The total erected cost of such an agitator with steel tank will approximate $2510. With an available depth of 1 1 ft. 6 in. and with pulp at a dilution of 1.5, this agitator will hold 134 tons of dry slime, making the total erected cost $18.73 per ton of dry slime capacity.
The Dorr Agitator 161
Fig. 101 is a view of this tank. In a flat-bottom steel tank is a central vertical cylinder or pipe carried by a shaft supported from the top of the tank and having two stirring arms with plows as in the Dorr thickener. The plows both agitate the pulp and draw it toward the center. The pulp is raised through the cylinder by means of air, supplied under pressure from an air pipe whose nozzle points upward at the foot of the vertical cylinder. The pulp, ri^g through the cylinder, is dehvered by two opposite launders attached to the cylinder by which it is distributed over the surface of the liquid contents. A continuous supply of fresh pulp enters at the intake at the left, and is distributed through the tank, and after thorough agita-
Fio. 101.— The Dorr Aititator.
tion and aSration. escapes by the outflow on the right. Thus we have a method of continuous agitation which has its advantages over agitatii^ in charges.
AGTTATIOn TRSATHBNT Dflution or ratio of solution to dry ore in the pulp undergoing treatment ia generally recognized as a factor to which proper attention must be paid if the highest extraction is to result. If a higher dilution is used than is necessary the capacity of the plant is cut down, and, on the other hand, if the dilution be too low the maximum extraction is not attained. It is quite unpossible to give general figures for minimum dilution as these figures vary for different ores. In general, lower dilutions can be used with gold ores than with silver ores, probably on account of the greater weight of metal to be dissolved in the latter case. Other conditions being equal, it appears that lower dilutions can be used with the Pachuca agitator than
with the mechanical agitator on account of the greater proportion (A air brought in intimate contact with the pulp.
The time of treatment depends upon the character of the silver and gold minerals and upon their d^ree of comminution, and, as previously pointed out, also upon the concentration or strength of the solution in cyanide. It therefore follows that finer grinding, or increase of the cyanide concentration of ihe solution, or both, will in general result in reducing the time of b«atment necessary. In cases where the cost of power is h^, and as a, consequence tiie cost of fine grinding would be exces^ve, the ore may be ground to the point where the minerals are liberated from the gangue and then separated into sand and slime, the slime being treated by agitation followed by either decantation or filtration. The sand containing the coarse mineral particles requiring a long period of contact for dissolution can be given the long period of contact necessary at a reasonable cost by leaching. Cases of this kind clearly indicate where combined sand and slime treatment can be employed to advantage.
Thickening the Slhne.— As a prelinunary adjunct to any filter operation, the pulp is first settled to a minimum moisture content; in average practice this will be found to approximate closely to 50 per cent, or equal parts of liquid and solid. In some plants this thickening is all done prior to dissolution, while in others, thickeners are used, both before and after the agitation.
For filtration, the moisture should be reduced to such a point that the heavier particles will remain in suspension, or at least settle very slowly during tlie cake-forming period. This governs the uniformity of tiie cake and is one of the most vital points in all filter operations, and since the fflze of the largest ;)article8 is, in tiun, governed by the limits of economic grindii^, the required buoyancy is best obtained by proper thickening.
The use of the Dorr continuous thickener has become almost imiversal for this work in America, and is being largely adopted in foreign countries as well. The machine requires a minimmn of power and attendance, and when used in conjunction with a diaphragm pump or air lift, for elevating, the discharge may be operated with practically no loss of mill head. Dis-chai^es as low as 33 per cent moisture are obtained in the Porcupine district, but on otJier ores careful attention may be needed to obt^ 60 per cent.
On the Rand, with colored labor at $0.75 or less per day, and power at $5.50 per H.P.-month, it is still found economical to retain the large intermittent settlers. These are steel tanks, varying from 50 ft. to 70 ft. in diameter, with from 10-ft. to 14-ft. sides and cone bottoms, givii^ an additional depth of from 4 to 8 ft. Peripheral overflows and adjustable decani ing arms are provided, and the tanks are emptied by sluicing the settled slime into the suction of a centrifugal transfer pump.
This system of settling or dewatering is particularly well adapted to African conditions, where flat open mill sites are used, and where all tanks
Agitation Treiatment 163
are unhoused. But it is to be noted that while the intennittent settlers on the Rand, particularly during the warm minuner months, frequently settle down to 40 per cent, and even to 38 per cent, moisture, the best that continuous thickeners seem able to do is 50 per cent. Since the extra 10 per cent <A water must be bnn^t up to treatment strength in cyanide, and then wasted, its eliminatifm is highly desirable. With any settlii^ equifsnent satisfactory moisture figures are seldom attuned during the winter months, when 60 per cent is more nearly the average figure, with, of course, the higher losses in cyanide and in gold.
The area required for proper continuous setting is, of course, mainly a function of the nature of the ore and the dilution of the pulp and varies from 4 to 15 sq. ft. per dry ton of daily capacity with a pulp feed oi from 90 per cent to 75 per cent moisture.
With intermittent settling as practiced on the Rand, the period required for dischaj^g introduces an additional time factor, and it is usual to allow 14 to 25 sq. ft. per ton of dry alive. It should be noted that, in this practice, practically all the water used in crushing and clasufication goes to the slime collectors, which thus handle a feed containing from 90 to 95 per cent moisture.
The cost of thickening operations will vary from $0,005 to $0.02 per ton milled, depending upon local conditions and the scale of operations.
Naturally, there is no standard size recommended or used, as this depends on the capacity desired, settling qualities of the pulp, density to which thickening is to be carried, clearness of overflow, alkalinity, temperature, and dilution of the feed. As a general rule, it may be stated that approximately 6 sq. ft. of tank area are needed per ton of granular slime per twenty-four hours, while 10 to 15 sq. ft. should be allowed for flocculent slime. Several installations in different localities show an averse of $2500 for the complete erected cost of a standard, steel, 30X12 ft. imit. In terms of tonnage of dry slime handled per twenty-four hours, the capacity of a 30X20-ft. standard thickener, under normal conditions, may be gjven as 125 tons of grStaular slune and 65 tons of flocculent slime, making the erected cost of the thickener $20 and $38.46, respectively, per ton of daily capacity.
Tbe Dorr Conttnuous Thickener.— As shown in F^. 102, this 30X 12-ft. tank has a slowly moving central vertical shaft with radial arms equipped with plows to bring the thickened settled material to its discbarge point at the center. The thick slime discharge is pumped to another tank for further beatment. Thb feed launder dehvers to a central drop pipe so as to cause no agitatitn. The clear solution escapes to the peripheral launder of the tank and thence overflows. The vertical shaft will revolve about once in twelve minutes.
Fig. 103, a sectional elevation of this thickener, shows its operation.
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The feed of this pulp at the center of the tank, the overflow of clear Uquid at the periphery and the discbarge of thickened pulp are continuous. There are four zones of settlement. At the top is a zone of clear wat«r A, beneath this-ia zone B, consisting of flocculated pulp of uniform consistency ;
Fia. 102. — ^Tbe Dorr Continuous Thickeoer. •
Fia. 103. — Dorr Thickener, Showing Slime-settling Zones.
directly beneath this is a transition zone C, and at the bottom a zone <rf pulp which is undergoing compression. The pulp in zones B and C is termed " tree settling."
CONTirrUdUS codnter-current decamtation This method of separating dissolved values from treated slime, by means of a series of Dorr continuous thickeners, is becoming very popular in
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Counter<:Urrent Decantation 165
America, paiticulajly in small plants, and more especially in thoee treating a granular product, which readily settles to 40 per cent moisture or lees and is amenable to treatment with very low cyanide strengths.
In operation, the slime passes through a series of tanks, the thick underflow of each being diluted with solution overflowing the second followii^ thickener of' the series. The soUds thus move constantly in one direction, while the solutions travel in the opposite direction. Water, in quantity sufficient to replace the moisture Anally discharged with the tailings, is added at the final thickener. The solution, traveling successively toward the head of the series and mixing with constantly richer pulp, is finally used in the crushing department, whence it overflows the first, or primary, settler and is sent to precipitation. The tanks are generally set so that sohitions gravitate throughout the series, and the necessary elevation of the thick underflow is made either with diaphragm pumps or with air lifts.
In most plants where continuous decantation has been successfully used, imderflows of 35 to 40 per cent moisture are usually maintained and solution equal to from four to six times the weight of the ore is clarified and precipitated. Under these conditions, the recovery of dissolved metals is excellent, but the loss of cyanide is higher than with filters.
This mechanical loss of cyanide is recognized as one of the principal factors limiting the use of continuous decantation without filters, where even moderately strong solutions are used. Solutions of less than ) lb. per ton KCy are seldom precipitated in American practice, and at this strength, the mechanical loss in the final residue will vary from J lb. to as high as 1 lb., depending on underflow moistures. In milling silver ores, there will be an additional loss, owing to the fact that dissolution of the metal continues as long as the slime is in contact with solution.
The actual operating cost of the thickeners is very low, and the sim- [dicity of the plant, enabling labor and supervisioD to be reduced to a minimum, must appeal strongly to operators of small plants.
As a most interesting adaptation of continuous decantation and vacuum filtration may be cited the practice at the recently enlarged Hollinger mill at Porcupine. Smtdl thickeners, operating as classifiers, separate the tube-mill product into amorphous and granular. The former is thickened and sent to a vacuum filter for washing, without other agitation than that obtained in grinding, pumping and thickening.
The granular portion is concentrated, thickened, agitated, and then sent to counter-current decantation tanks for washing. This system eliminates amorphous or colloidal material from the bulk of the tonnage, where a very thick underflow is imperative, and, at the same time, furnishes a satisfactory product for the filter.
This is now known as the C. 0. D. system. The flow sheet of it, shown in Fig. 104, illustrates the method where there are four Dorr thickeners.
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Cyaniding Of Gold Ores
W, X, Y and Z in series. It is assumed that the crushing is done in cyanide solution, the overflow from the thickenii^ Tank X, being returned to the mill for mill solution. The ground pulp enters tank W, whose overflow, called thepregnantsolution, goes to the next step in cyaniding, the precipitating of the gold frton the solution. After having here deposited its gold content the solution, now called " barren," is used to dilute the underflow of thickener X, as it enters tank Y. The overflow of thickener Z is also mixed into the feed to Y, receiving also water for washing ita pulp, which a then sent to waste. The overflow from Y meantime enters X, together with the partially exhausted thickened pulp from W. It is thus seen that pulp passes from tank to tank from left to right, losing more and more of its gold, to be discharged and exhausted from Z; while the wash-water flowing into Z picks up an increasing load of gold as it encounters the progressively a I richer pulp in its passage to the left. It fi-
I N, nally leaves Wat its full possible strength.
This method is best adapted to low- t ^o^fagr.y^ grade, easily leached and settled ores, and
I w\y/ it dispenses with agitation other than that
resulting from the crushing and clas^y- ing and the movement through the thickeners, and likewise with filtration about to be described.
At the Hollinger milt, Porcupine District, Ontario, Canada, where some 50,000 tons of a soft quartz ore with schist and pyrite is treated monthly, there are five sets of 40-ft. tanks as just described. ITie tanks are array^ed with a difference of elevation of 2J ft, between the steps, the last of the series Z, being the highest so the solution flows from tank to tank and to precipitation. Thence it is pumped to the mill-solution feed-tank. The cost of Fio. 104.— Continuous Couuter^^ui'- decanting is given at 2.09 cents and the rent Decaatatioo. taiUi^ cany 9.75 cents per ton.
FILTRATION OR SEPARATION OF UETAL-BEARHTG SOLDTTON FROM SLOfE RESIDUE
Probably the most vital point in the practical application of the cyanide process is the filtration or separation, after dissolution, of the metal-bearing solution from the slime residue. Certainly no other point has called forth such a combination of inventive ingenuity and practical ability
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Slime Filtration 167
as has been expended in developing a satisfactory technical and economical BohitioD of this problem.
The reason for this is readily appreciated when one stops to conuder that for every unit of dissolved metal finally discharged witfn the reddue, there is incurred a net loss equal to the market value of the unit phis the mechanical loss of cyanide, amounting to a further $0.05 to t0.10 per ton. Since filtration or decantation is almost the last step in gold and silver production, and amounts at most to 5 per cent of the total cost, there is obviously every incentive to obtun the highest possible efficiency.
The amount and character of this material produced will depraul chiefly upon the nature of the ore and upon the d^ree of comminution necessary to obtain an economic extraction. Many Bo-called " aU-slime " plants find it feasible to agitate and filter, or decant, a product of which fully 40 per cent will remain upon a 200-mesh screen, while others grind to a point where only a fraction of 1 per cent will remain upon this mesh. The governing factors are purely individual, such as size and nat\u« of plant, location, character of ore, etc., and can hardly be generalized.
All ores after fine grinding may be classified into two products — granular and amorphous, and most of the difficulties experienced in shme filtratitm and decantation may be traced to extreme conditions as regards either the one or the other.
An undue amount of coarse, granular material will not only increase the power consumption in the agitators and thickeners, but will also choke and cause frequent interruptions, and unless filtered at 50 per cent moisture, or less, will result in classification and uneven washing.
On the other hand, where the percentage of the amorphous product is high, it is seldom possible to thicken to less than 60 per cent moisture, even with a very large settling area, and a pulp of this dilution will not give economic results with either continuous or intermittent decantation, and even for filter treatment requires a largely increased area and frequently results in slow and imperfect washii^. Cracks and channels will also develop with any type of filter in which the cake is exposed to the air before or during washing.
It may be safely stated that for ideal slime-washing results with any of the methods now in vogue, an all-slime product should fulfill the following conditions.
It should be ground in closed circuit with suitable clasdfiers so that not more than 25 per cent of the total product will remain upon 200-mesh screen and all of the metallic or sulphide portion will pass the same aperture.
It should settle from 85 per cent moisture to at least 50 per cent with a continuous settling area of 6 to 10 sq. ft. per ton of dry solids per day.
The various processes and machines now in use include 95 per cent of the slime tonnage produced in the cyanide process throughout the worid.
Main Systems Of Filtebino
1. Thickeoiiig. — The Dorr continuous filter; settling tanks.
2. Vacuum FUtratioii. — Butters, leaf ; Moore, Oliver, drum; Portland, drum; Ridgway, leaf; American, continuous suction.
3. Pressure Filtration. — Merrill sluicing plate-and-frame press; Dehne, plate-and-frame-presB; Kelly, enclosed-leaf; Burt, revolving cylinder; Sweetland.
^ 4. Continuous Decantation.— Dorr eystem. 5. Intenpittent Decantation. — Baud EFystem.
Fio. 105— The ButUr Vacuum Filter. (3) VACUUM FILTRATION The Butters Vacuum-leaf Filter. — Fig. 106 is an elevation and Fig. 107 a view of a filter leaf on which a layer of slime haa been built up and part of it removed to show its thickness. A vacuum filter plant consists of several filter tanks (see Fig. 105) with vertical sides and a V-shaped bottom, in which are suspended a number of filter-leaves constructed of pipe, with either cocoanut^matting, wooden laths, or ripple iron as a support for a cloth which is sewn around and covers the whole as shown in Fig. 106. Leaves vary in construction and 5 by 9 ft. is a handy size. Connected with the vats is suitable piping and centrifugal pump of large capacity for filling and emptying with slime, wash-solution, or water as desired. In operation, the filter vat is filled with pulp
Vacuum Filtration
to s point over the top of the filter leaves, and a valve opened connecting the vacuum pump directly with the Sltets. Clear cyanide solution is drawn from within the filter leaves and delivered to a clarifying tank for precipitation, the sUme remaining aa a cake on the outside of the filters. The filters are kept submerged by refilling the vat at intervals, and jets of air are introduced at the points of each hopper to keep the slime in mispenHion. When a cake I in. to IJ in. thick has been formed, the surplus pulp is pumped back to the stock pulp vat, and the box is then filled with solution to wash the cake. During the time of forming and washing the cake, the vacuum is mainlined at the highest possible point, but when the cake is exposed to the air during the transfer of pulp and wash solution the vacuum is reduced to 5 in, to prevent the cake cracking. It is possible to form a' cake with some ores in five minutes, while others take up to ninety minutes.
Sufficient wash solution, about 2 tons per ton of slime, is drawn through the cake by the vacuum pump to effect a complete displacement of the original valuable solution. A portion of ihis solution goes to the clarifying
tank before precipitation. When the wash is complete, the v connected, and a flow of solution or air is introduced into the interior of the filters, causing the cakes to drop. The mass of thick sludge remaining in the vat is diluted with water and agitated with air for a few minutes to make a homogeneous pulp of 1 to 1, which is then pumped to the residue pond. Typical cycles of operation are as follows ;
—Butter's Filter-frame.
R«ml OiB,
45
15
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The Oliver continuous revolving filter ae shown in Fig. 108 consiBts of a drum or cylinder with open ends, rotating on a horizontal axiB, with the lower portion submei^ed in a tank containing the pulp to be filtered. The surface of the drum is minded into compartment or sections, the
Fia. 107.— Slime-cake partly removed to show eecticai.
Pia. 108.— OtivCT Fnier. Fia. 109.— The Amerimn Filter.
divisions running parallel to the shaft. These sections are covered by a screen, and a filter medium is stretched over it, being held in place and protected from wear by a wire winding. Each scclion of the dnun is connected by two pipes passing through a hollow trunnion to an automatic valve which controls the application of the vacuum for forming and waah-
Suction Filtration
ing the c^e, and the admisfiioii of air for its dischai^. A ecraper is fitted across the tank and rests on the wire winding 90 that the wa^ed cake is removed when released by air. Vacuum hltration costs vary from 5 to 10 cents per ton.
The American Continuous Suction Filter. — Fig. 109 shows a view of a three-disk filter and Fig. 110 an end view. Each disk is made up of eight s^ments. At one end of the shaft is a distributing valve having eight openings, each to one of the eight segments. The valve housing has three inner recessed ports. Port No. 1 on the underside connects to the filtrate suction line and applies suction to the four submerged 1eaf-eegment«. Port No. 2 connects with the wash-water suction and drying line, taking
Fio. 110. — EdJ View of "American" Filter.
care of the three upper left-hand leaf s^ments. Port No. 3 admite compressed air to a segment as it passes this port. The air is admitted for a few seconds only, inflating the filter cloth of the segment and loosening the formed cake thereon, while the scrapers remove it so that the product drops into a hopper or conveyor below.
Both the Kelly and the Dehne presses have been extensively used. They have the advantage over the suction filters that high pressure can be carried, since against 10 lb. suction in the one we can cany 50 lb. ot more in the pressure filter. This makes for quick filtering and washing,
Dehne presses are used extensively for concentrate in Western Australia and at Waihi, New Zealand. Three-in. cakes are formed At 40-lb. pressure and are washed satisfactorily.
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The Kelly FQter. — Fig. Ill ^owa a twin unit of a filter, the right-hand unit open for discharginf; the washed precipitate. Into the closed unit at the left the pulp is piimped filling the cylinder around the filter leaves as seen at the rlghl^ hand unit. The escaping solution through pipes that connect to -each leaf is discharged to a launder. This continues until a thick cake has been built up on the leaves. The excess of unfiltered slime is run out and fresh water is introduced to wash the ■ caked accumulation. To dis- i charge, the front head of the
I
9 b unlocked and the head with the attached carriage which ". sustains the filter leaves is run (2 out, as shown in the right-hand unit, this unit by the same movement being closed and locked. The cakes are dislodged by inflating the filter leaves with air. The carriage is run into the cylinder and the head again automatically locked in place.
The Kelly press is used at the Goldfield Consolidated and the Alaska Treadwell with satisfactory results. The Rectangular or Dehne Filter Piess. — This, as shown in Fig. 113, consists of a heavy frame carrying forty filter leaves 42 in., square, one standing at the left leaning against the press and forty filter frames, one of
Pressure Filtration
FiQ. lU.— Sweetland Filter Press.
Flo. 115.— Merrifl FSter-press Installation.
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Cyaniding Of Gold Ores
these showing at the right. The filter leaves are covered by a canvas filter
cloth, one on each side. Fig. 116A shows the roughened surface of a leaf, or
plate, the surface grooved so that the filtered 9oluti<m is carried by the
grooves to the outlet channels, circular opening^ at the sides of frames and
leaves. The filter cloths are pierced
with openings at the channels for
the circulation of the pulp and the
filtrate. These cloths are huge
enough to extend beyond the
frames. The follower, a solid plate
at the right, is now brought against
the assembled parts and the joints
tightly compressed by the follower
screw.
Filling is generally done by a
three-throw pump, which will
charge a press, forming a 2-in. cake FlO. lis.— Merrill PreaB-frame. . , '^ . ,' . .,.° . ,„
in from eight to fifteen nunutes according to the thickness of the pulp from the last a^tating vat. The final pressure is as much as 60 lb. per square inch. In operation, ttie pulp flowing along the left-hand channels of the frame finding no outlet, the solution passes through the filter cloths, leaving the solids behind. The solution from each leaf flows out through a cock at the left-hand lower comer into a launder set beneath and thence to the gold solution tanks. Washing the cakes is the next operation, taking thirty minutes at 75 lb. pressure per square inch. The wash-solution goes to
the wash-solution vat«. _ ,,„, ., „, „ ,^
, ... FiQ. 116A.— Merrill Presa-plate.
The cakes are now dned
by compressed eii for about two minutes, the press is opened and discharged. A cycle of operations takes seventy-two minutes.
The Merrill press was the outcome of treating low^^rade slime, and at the Homestake this is filled into the presses, the gold dissolved in them and the reffldue automatically discharged without the apparatus being opened, except for necessary repairs. It has been very successful at this mine and
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Pressure Filtration 175
otheiB in North America. In principle ihe Merrill press is similar in many tespects to the Dehne, The Merrill is essentially of the ordinary rectangular fluBb-fdate-aiid-distance-frame pattern with internal channels, but equipped 'with the automatic diachargng device which is the distinguishing feature of the press. A standard fweas to hold 25 tons of alive contains up to ninety-two frames 4 in. thick, each with a cross-sectional area of 25 sq. ft. Between the f nuues are the usual solutitnt plates, and all aie made with suitable channels for flow of solutifxis. When empty, a press wei^ about 70 tons. The filling and washing is similar to the Dehne, only pressures are about half of that used in the latter type. Slime may be leached in the preas or not, according to tits character and vahie of the ore. In discharpng a Merrill press the procedure is as follows : In a lower centrd channel in the frames is a 3-in. pipe, restmg on supports, bolted to plates at intervals, throughout Ifce length of the press and connecting at the front standard with the water supply and rotating mechanism for the [Hpe. Projecting into each frame compartment from the sluicing pipe is a 0.16 in. noble through which water ia discharged against the cake ol slime, while the whole pipe is rotated throu|^ an arc of about 200°. The cycle <rf operations at the Homestake mill is 560 minutes and at Santa Gertrudiq mill ninety minutes; at the former the slime is leached in the press and at the latter it is previously agitated.
As typical of the operating cost of a large Merrill filter installation in Mexico, may be taken the following figures from the Esperanza Mining CcHnpany at El Oro. Approximately 1000 tons of slime are filtered daily with six presses, averaging eighty-two frames each, being equivalent to 100 lb. per square foot per day. Caking effluent carriefl t3.20 in gold and 1 oz. silver. Dissolved metal loss is $0.03 in gold and 0.01 oz. silver. Operating charges are: Canvas, t0.0108; acid, 0.0035; labor, 0.02; miscellaneous, 0.00005; slmcing water, O.OlffiZ; a total of S0.045 per ton filtered. The item for sluicing water represents all charges incidental to settling and returning for re-use all water s^it out with sluiced residues.
As illustrating the additional dissolution of metal which almost invariably occurs during washing in pressure filters of this type, the following data from the Merrill installation at the mill of the San Luis Mining Company, in Durango, may be cited : The caking effluent carries 25 os. silver per ton, while re-washed filter beads and tails show 4.7 and 4.08 oz., respectively.
General Remarks On Filters
For large output are used the various types of suction filters, whether intermittent like the Butters, or continuous like the American or the Oliver. Where, however, the temperature of the solution approaches the boiling-point they are not effective, due to the vapor generated imder vacuum.
Their highest suction is not to exceed 12 lb, per sq, in., but this may be an advantage due to the fact that the lower the suction the more open tJie porous structure. Colloidal slime is very apt to pack, makii^ slow filtering. One may note that in the American suction filter the mud layer is automatically washed in the revolution and is scraped off against an inflated filter cloth at the critical instant.
The filter presses, such as the Merrill or the Kelly, work under a high pressure of 40 to 50 lb. t« the square inch, and so act rapidly, especially on granular material. They work well on hot solutions. In the case of tiie Kelly press the frames must be washed, then withdrawn and \mloaded. Sometimes the preas~men get careless and do not wash the cake; this results in loss of valuable solution. With less slime proportionately to handle they are at their best, since stoppages to unload are less frequent. The Merrill press has the advantage that it ma^ be unloaded ready for the formation of a fresh cake without having to open up. The press is well adapted to clarification or to filtering precipitate from cyanide solution, since the amount filtered out is small compared with the total bulk of the Bolutian. Washing and air drying can be well done in the press.
Clarifying
The solution from decantation or from filter pressing is not always quite clear, A little slime may get into it by Icak^e through the filter cloth, or an overflow may at times be turbid. Such solution may be passed through a sand filter, vacuum filter, or a filter press of a few leaves; with little solid matter this is quit« readily done, and leaves a bright and sparkUng liquid. If the solution is not quite clear a little slime may deposit on the precipitant, whether zinc shaving or zinc dust, thus diminishing the activity of precipitation.
Of the Alter presses the Merrill occupies a small floor space, has a capacity of 1000 tons of solution daily and is readily cleaned without openii^ the frames. To this end the press is sluiced out every six hours, using barren solution. The filter cloths are of No. 10 duck, and last six weeks. They get coated with a lime and alumina deposit, so that they must receive a HCl acid treatment about every three days. To do this an O.d per cent acid solution is pumped through the press at a pressure of 50 lb. for an haai. This is withdrawn, and stored for re-use, while the press receives a water wash. The slime caught by the press is sent back to one of the thickeners, since it still has gold values and cyanide.
' Where sand filters are in use an extra tank is required in reserve. A sand-filter plant of ample capacity occupies a large floor space and is ejtpensive to erect.
Unquestionably, the handUng of slime, as produced from the operation
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Crowe Vacuum Process 177
of the cyanide proceae, has reached a point where no further radical changes should be expected. Tbe operation of separating the dissolved metal from the treated residue may now be carried out at a cost of S0.05 per ton, imder favorable conditione, and recoveries of 98 and 99 per cent are not at all unusual. Themechanicsof the various machines will, of courae, be modified and improved to suit the needs of special problems, and there is always the possibility of reducing capital cost. In comparison, however, with the development of the last ten years, these further modifications will be relatively unimporiAnt.
The Crowe Vacuum Process
This consists in removing from the solution, just prior to precipitation, substantially all the dissolved air and oxygen. Generally, cyanide solution going to precipitation is saturated with air absorbed during the air agitation of the slime pulp, this air containing 30 per cent of oxygen and 65 per cent of nitrogen as against 23 per cent of ox^en and 75 per cent of nitrogen as found in ordinary air. Indeed in zinc boxes exposed to cold in winter a white precipitate, a hydrated zinc oxide, will form, due, doubtless, to the dissolved oxygen of the solution. In presence of the zinc precipitant the oxygen polarizes or reverses the action of the precipitating couples and causes re-solution of the precipitated gold (or silver). Complete precipitation can take place only when sufficient hydrogen has been evolved to combine with the free oxygen, and in this evolution results the consumption of both zinc and of cyanide.
The Crowe apparatus consists of a receiver or drum 4 ft. diameter by 10 ft. high. The solution, coming from a steady-head tank set 24 ft. above the top of the receiver, pours through the top downward over a series of perforated trays. This breaks it up into a spray. A vacuum pump continually sucks away the occluded air of the solution. The solution level at the bottom of the receiver ia maintained .30 in. deep by a float operating a butterfly valve in the intake pipe. The solution flows away at the very bottom to a pump set 33 ft. below so that one may be sure that the pump foot-valves are covered and that no air can be drawn in when pumping. It is into the suction pipe of this pump that the zinc-dust is introduced for precipitating. It is computed that by using this process, a saving of 50 per cent of the zinc dust can be effected.
The Precipitatioh Of Gold From Ctakidb Soutfiohs
Precipitation. — We have our choice between the ordinary extractor-box, using zinc-shaving, and the Merrill precipitation press, using zinc-dust.
Zinc-boxes are efficient, but occupy considerable floor-space, and are not simple when it comes to the clean-up. There is always the aggravating problem of the " ainc shorts."
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Merr^ presses are compact, thief-and fireproof, and allow a quick clear-up. In precipitating strong (0.2 to 0.3 per cent KCN) and rich solution with concentrate treatment, they ^ve good satisfaction.
The Bibbull Preopitatioit Process
This consists of the introduction of zincKlust into the suction of a pump from a pregnant solution tank, or the flow from the Crowe Vacuum treatment, by means of a special feeder thitn^ a filter press. Fig. 117 shows
Fia. 117. — Merrill Precipibttion Appantua.
the preferred type of feeder for zinc dust. This is contained in a hopper marked " precipitant feeder." By means of a screw feed the dust is fed in an accurately adjusted manner through a small hopper into the " precipitant mixing cone " there mixing with a stream of solution. The mix-
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Merrill Precipitation Process 179
ture is maintaioed at the proper height by means of a float-controlled valve at the bottom of the cone.
The design and construction of a satisfactory feeder is considerably more difficult than might appear at first sight. The original type as proposed by Merrill was an endless belt. On the top surface of this belt was placed the amount of zinc necessary to precipitate the given tank of solution. The belt was actuated by means of a series of floats, so as to cause the belt to travel forward and feed the sine thist in exact proportion to the rate at which the eolutim was lowered in the sump LaniEs. A later type, illustrated in Fig. 117 makes use of a screw located at the bottom tA a hopper, the hopper being provided with a hanmiering arrangement or a reciprocating arm to prevoit arching of the zinc dust. Provision fturegulating the speed of the screw is made by means of double cone and beH drive.
Operation. — ^The rate of zinc feeding is checked by weighing the amount run in in five minutes, experience deciding how much will be needed according to the assay value Of the solution. When a newly cleaned press is " cut in," that is, brought into use, 20 lb. of zinc-dust is added at once, and for six hoiu« the rate of feed is doubled, all to form a zinc^ust coating upon the filter cloths of the press. Also the solution for the first fifteen minutes is returned to the gold or pregnant solution tank as having been imperfectly precipitated. Where two presses are operated the flow is turned to the first one until it is normal, then the flow is throu^ both.
The Ctecn-op. — The press is dressed with four thicknesses of cotton sheeting, the outside cloth being removed at each cleaning (every five or six days) and a new one added on the bottom. To clean a press, the solution is cut off, the press drained, then blown with compressed air for sixty to ninety minutes; this dries the precipitate to about 45 per cent moisture. The press is opened and the precipitate, amounting to perhaps 130 lb. is scraped into the precipitate wagon set beneath. The outside cloths are burned, the resulting ashes being added to the rest.
The barren solution from the Merrill presses is returned to No. 4 thickener.
The method of piec^tation enqik^ed, assuming equal efficiency as regards the precipitation and recovery from solution of the precious metals, may exert an important influence upon extraction, either through introduction of the precipitant used into solution or its failure to precipitate certain interfering elements. For example, zinc in the presence of arsenic may interfere in the treatment of certain ores. If aluminum precipitation is used the difficulty ia overcome through the elimination of zinc. When copper reaches a certaia concentration in solution difficulties arise with both extraction and precipitatitm. Neither zinc nor aluminium precipitate
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Cyaniding Of Gold Ores
copper to any extent, hence if copper is to be removed from solution electrolytic precipitation must be used.
Carbon, which at times occurs in gold and silver ores, may occaeioa difficiilty in cyanidation. It has been generally assumed that carbon occurs in gold and silver ores in the form of graphite, but the evidence available by no means support this view in aU cases. The two extremes of carbon as regards its behavior in cyanide solutions are graphite and charcoal. Graphite ia dense and doee not poaaess pores, therefore cannot occlude gases, while charcoal is porous and has the property of occluding relatively large volumes <>f gasee. Graphite doee not precipitate gold and silver from cyanide solutiona while charcoal does. Intermediate between these two extremes are various forms of carbon which will precipitate gold and sUver to a greater or lesa extent.
Feldtnumn has shown that the graphitic or carbonaceoua schist from the mines of West Africa will precipitate gold from cyanide solutiona roughly in proportion to the carbon content of the schist. Gold so precipitated is not soluble to any appreciable extent in fresh cyanide solution, but is soluble in sodium sulphide solution. He thinks that the compotud formed is possibly carbonyl aurocyanide, which may react with sodium sulphide accon^g to the following equation :
However, the graphite was found definitely to interfere with cyanidation. Experiments witb deflocculation of the colloidal portion of the ore including the graphite, gave negative results, but it was discovered that if the physical state of the graphite were altered, most conveniently by heating, the extraction was wonderfully improved. Hie results below ahow clearly the effect of heating:
Table I
P«CBDt
o„ft,„.
U.60 00.10 90.10
Wet slime boiled with w«t«r, graphite diiiumed ofT. Wet sliiae boiled with mt«r, Eraphite not skiinmed
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Precipitation Methods 181
Attempts have been often made, with some degree of suoceaa, to precipitate the gold electrolytically upon platee suspended in the turbid solu-tioD. In present-day practice, however, the solution is invariably clarified before precipitation is attempted. The solutions from sand-leaching are generaDy clear enough for precipitation, but the solutions fnun slime ^treatmeait need clarifying after filtration.
When the slime and sand treatments are combined much of the solution from ihe slime treatment can be clarified and built up in value by using it for the first washes upon the sand, but even then the solutions from the sand plant should be later clarified before attempting precipitation. This has been done by the use of the sand filter, but this is of limited capacity, and the preferred method is the use of a leaf filter of the necessary raze, as already described under head of clarifying.
An expedient sometimes used has been to fill the head compartments of the zinc box with excelsior or other fibrous material to act as a filter. This must be removed at frequent intervals and washed. However, at present the Merrill clarifying press is preferred.
Gold dissolved by cyanide solutions is recovered by pasrang the solutions through zinc shaving, zinc dust, zinc wafers, aluminum dust, or charcoal. The last precipitant is not used much, as, although efficient, it is rather a nuisance in requirii^ a good deal of attention during operation and at cleanup. The most commonly used precipitant is zinc-shaving arranged is a long narrow box, in which the solution is made to flow up through the mass. This is the original MacArthur-Forrest process, llie reactions involved are fairly well underetood, and in depositing gold the eqiiation is as follows:
(13) KAuCNa-h^CN-l-Zn-|-HaO = K2ZnCN4-|-Au-t-H+KOH Generally speaking, gold-bearing solutions should be do lower than 0.03 per cent KCN strength for good results, and if lower, a drip of strong solution, or lumps of KCN added at the head of the boxes will strengthen ^em. If copper has been dissolved from an ore, it will precipitate on ihe zinc, thus preventing proper deposition of gold and silver. A partial remedy for this is to either dip the shaving in a 10 per cent solution of lead acetate.'w add the latter solution regularly at the head of the boxes. In fact, a little lead acetate is at all times quite useful in the boxes. Also the addition of strong sohition will delay precipitation of copper. Silver ores always make more precipitate than gold.
THE Zinc OR EXTRACTOR BOX
A description of a typical apparatus used in precipitation is as follows: A iaac box (Fig. 118) contams seven comparbnent«, each 12X15X24 in. These have perforated false-bottonu of sheet-iron or wir&-cloth that bold
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Iss Cyanidinq Of Gold Ores
up the nnc-shaving with vhicb they are filled. The partititHis are set altemateiy up and down, to cunpe) an upward Sow of the gold-bearing aohition thnm^ ihe shaving, and to bring it intimately in contact with and iofflire the precipHatioQ of the gold upon the surface of the zinc. At a in the side elevation the solution enters the box through a pipe m at the left, passed through all the compartment, flows over the last partition, and dis-charts through a down-turned pipe into the sump tank. The box is set at a grade of ^ in. to the foot.
CROSS SEcnoN
well-deaigned Zino-boi.
In Fig. 119 we have a section of a well-designed zinc box and launder <rf sheet steel with wooden interior partitions. The supports of three screens (forming the false bottom through which the solution rises) is as indicated. The bottom is inclined to drain to an outlet which itself is plugged from the in^de.
The Clean-up of the Zinc-btnes. — This is made monthly or bimonthly, according to the bulk of the precipitate to be treated, and tiie need of realizing values for operating expenses.
CleaiMng ttie Zinc Boxes. — ^At the time of the clean-up, only one zinobox is taken care of at a time, the flow continuing in the otben. The
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Zinc-Box Clean-Up 183
flow of gold solution to this box is stopped, and water is run in to di^lsce the solution contained. Banning in the first compartment, the fine-material is lifted out, while the Bbaving is agitated in the water with the hands protected by rubber gloves. This ia not done roughly, for the brittle shaving would be imnecessarily broken and the water would be black with the fioating precipitate. The plug in the side (see k in the croeft-eectitm Fig. 119) is gradually withdrawn, and the accimiulated slime and water allowed to flow into the launder h. The plug is replaced and the compartment is again filled with water. The zinc again is rinsed and rubbed, and the loosened precipitate once more drawn ofE. About three such washes free the shaving from precipitate and short-ainc. The comfiartments are thus successively cleaned up, and the shaving from each compartment is moved toward the head, and in the last compartment, where needed, replaced by fresh shaving. Finally the launder is cleaned with a hose, and everything washed into an acid-tank or clean-up sump along with the material first lifted out of the compartments, and sulphuric acid added.
When all the boxes have been cleaned, the precipitate is allowed to settle a short time in the acid-vat and the supernatant liquid is siphoned into a settling tank. In this lai^r vat the particles of precipitate have opportunity to settle, and to be recovered subsequently in the filter-press. The acid-vat is stirred by hand with a wooden hoe, or preferably a power-dnven agitator in constant motion. This insures a thorough agitation of the sludge and precipitate in the acid treatment.
The Add Treatment — Upon the watery slime about 30 lb. of sulphuric acid is poured. This acts upon the short zinc and produces a violent effervescence. After subsidence the whole is stirred. When the action again abates 15 lb. of acid and the same amount of hot water are added with occaaional stirring. This is repeated until further addition of acid produces little effervescence. Then the mixture is allowed (o stand two hours, and a portion is tested with more acid to see that decomposition is complete. The total time for the operation is foiuto six hours.
Filter-praasing the Precipitate.— The black mixture, containing zinc sulphate in solution, is diluted with hot water to within a few inches of the top of the vat. The whole content is stirred and then pumped through a lead-lined filter-press, see Fig. 1 13. The vat is washed and the washings are also pumped through the press. Finally the residue in the press is washed with hot water to entirely remove the sine sulphate.
The entire precipitate, having been trmisferred to the press, while in this position is washed with water under pressure. The water is followed hy ccnnpressed air to dry the precipitate, which, after this, is ready to discharge. To dischai^ the press, the tightening-screw is slackened, the
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follower ia drawn back, and the frames are aucceseively separated. The 'grayish black residiie in the recesses of the frame containing 20 per cent or less of water, drops into a drying-pan placed beneath the press to receive it.
Dressing the Boxes. — At the start, compartments Nob. 1 to 4 inclusive are packed with fresh zinc shavings; in two days the fifth and in two days more the sixth compartment. As zinc is consumed fresh shavings are added and this dressing of the boxes is more frequent as the time of the monthly clean-up approaches. The first compartment, however, takes already-plated long filaments from the next compartment and on top of them shavings now nearly consumed, the so-called short zinc,
DRYING AITD REFINIirG THE GOLD PRECIPITATE
The product, still damp, is transferred to pans 24X44 in. by 4 in. deep. A pan is slid into a cast-iron muffle, and heated until the precipitate is dry and finally to an incipient red. It is then removed, allowed to cool, and the weighed contents cautiously mixed with 50 per cent borax, some sand and soda. Since the product is light and dusty, care must be taken in handling it, and for fusing it must be put carefully into melting furnaces. The molten metal is stirred in the crucible, then poured into crucible molds, and o^ cooling, the slag is removed and the gold remelted into an ingot,
SEPmiHG WITH BICHROHATE
Where a filter press is not used a good method is to proceed as follows: To the acid-treated and washed precipitate is added five times its estimated weight of water, altw sulphuric acid and bichromate of potash, but each stirred in separately in the proportion of four parts of 60° acid to one part of the solid bichromate, the latter first dissolved in hot water. Careful additions arc made until a slight coloration, still showing, tells that enough bichromate has been added. The remaining precipitate is decanted, well washed, and dried. It is transferred to a clay crucible with an addition of borax for fluxing and niter for oxidizing carbonaceous matter, is melted and poured into an ingot free from zinc, lead, or copper.
Refining in a Cupelling Furnace.— As practiced at the Alaska- Treadwell cyanide plant, the acid-treated precipitete, containing 17 per cent of silver, 5 per cent lead, 14 per cent copper, 5 per cent zinc and 20 per cent insoluble matter, is treated on the hearth or test of an English cupelling furnace (see Fig. 278). The furnace is charged hourly with the precipitate to which have been added fluxes. This melts down into the lead bath contained on the hearth. Here it is oxidized, yielding slag and
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Refining Gold Precipitate 185
lead-bullion, which are tapped off intermittently at opposite sides of the hearth as they accumulate. A typical charge mixture would consist of precipitate 100 lb., glass 22 lb., sodium carbonate 25 lb., old slag SO lb., iron tumings IS lb. Such a charge would yield about 35 lb. of rich lead, 10 to 15 lb. matte, and 160 to 180 lb. of slag.
Diying and Refining at the United Eastern Cjnmlde Plant — After determining the moisture content, the undried raw precipitate is fluxed with 11 per cent borax glass, 11) per cent sodium bicarbonate, 6 per cent manganese dioxide, 3.3 per cent ground bottle glass, and at least 10 per cent of old slag shells from former melts, the percentage being in terms of the calculated wei^t of the dry precipitate.
A precipitate press ordinarily runs from five to six days, and yields about 130 lb. dry precipitate. In resuming precipitation after a final clean-up, one press is given the entire flow. When this builds up a pressure of 35 to 40 lb., the second press is opened just enough to maintain the pressure of the first press below 46 lb. When the second press reaches 20 lb. pressiue, the entire flow is turned into it wid the fitst press is cleaned. This method is carried on until the end c^ the month, when a final clean-up is made. The solution is metered by a revt^tion counter on the triplex pump, which is calibrated at intervals withalcnown tonnage of solution. Solution samples are taken each shift; the heads are a dip sample every hour and the tails a drip sample from the barren flow. The tonn^^ and the average solution oBBays give the ounces of gold precipitated daily; this is checked monthly against the bullion sold.
Melting. — The fluxed wet precipitate is placed in No. 5 paper bags and fed to an oil-fired No. 150 Case tilting furnace, using a No. 100 long-lipped, graphite pot. When ready for pouring, the pot contains fifteen sacks of precipitnte and yields a 600-or 700k>z. button and some 40 to 50 lb. of slag. Thb chaige is poured into a conical mold and allowed to set a few minutes. The slag is then tapped through a hole about 2 in. above the gold button, and run into cold water for granulation. The cold skull of shell left in the mold, which contains most of the shot, is put back with a subsequent charge. The granulated slag, which carries some 25 oz. of gold per ton and as much silver is ground in a small ball mill and concentrated with a laboratory-sized Diestcr table. The resulting slag tails from each month's run, of which we usually have less than 400 lb. carries a total value of about $50. This is shipped to the smelter once a year.
An average month's run will show: Crude dry precipitate, 21,000 oz. troy or 1440 lb. avoirdupois; 71 per cent bullion yield, 14,936 oz. troy or '1024 lb. avoirdupois; 39.7 per cent gold yield, 8538 oz. troy or 573 lb. avoirdupois; 21.4 per cent silver yield, 4497 oz. troy or 308 lb. avoirdupois.
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The crude ineeipitate ccmtains from 6 per cent to 7 per cent of zinc and about 10 per cent of lead, the latter comma frcnn a lead acetate drip added as the, solution leaves the clarifying fitter.
The bullion buttons are temelted and cast into bars wei^iing about 150 lb. each. A dip sample is taken mih a lO-^^m. clay crucible just before pouring. This sample is granulated in cold water and is sent to the assayer. The bullion, as shipped, has an approximate fineness of 560 in gold and 301 in silver.
Caittal Costs Of Slihe Plants
It may be of value to cite approximate figures from Rand practice, vdiere the intennittent decantation and pump tranter type of plant has reached such high develo[»nent. The tetal approximate ciqiital costs of slime plants, in dollarB per ton of slime treated per twenty-four hours is as follows: 150 to 225 tons, |292; 300 to 450 tons, $276; 600 to 900 terns, $260; 1200 to 1800 tons, $243. Such slime plants provide for collecting and two washes, two days being available for collecting and treatment, and include complete pulp, solution, decanting, and water eervioes, t<^ther with solution clarifiers, but exclude precipitation, refining, and power plants. It mxist be noted that Rand plante require very little expenditure for buildings.
Costs Of Dissolution B7 Sldcb Agitation
Coste of dissolution may be of some interest, although it must be remembered that such coste depend on numerous widely varying conditions, such as character and grade of ore, cost of labor, power, and supplies, amount of extraction effected prior to agitation, arrangement and capacity of the plant, and the recovery percentage achieved. Furthermore, such accessory steps in the cyanide process as fine grinding, filtration, or decantation, precipitation, melting, and refining, will depend on the dissolution method in use and, hence, should be taken into account in comparing " slime fetation " with " sand leaching " or " filter treatment." Then too, tiie higher percentage of dissolution obtuned by " slime fetation " must be weighed against the value of the dissolved metals and cyanide lost by imperfect filtration or decantation. It is to be understood that the foUowii^ figures are merely numerical averages of compilations from representative plante in various parts of the world. The cost of dissolution per ton of sHrne is made up of the charges for thickening, f^tating, proportionate pumping, supervisioD, cyanide, lime, and lead salte.
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Filtration Vs. Decantation Cost Op Filtration Or Decantation
I0.26 0.42
Included in the above, but so far below tbe average as to require special notice, is the intermittent decantation method practiced on the Band, where the Blime is particularly amenable, in so f ar afl concerns thickening, dissohiiJon, and chemical consumption. Figures from nine Rand slime-plants sbow an average cost of dissohttitm, decantation, precipitation, and refining, of 24.5 cents per ton of slime, the extraction averaging 85.9 per cent on slime heads of S1.92. Recent Rand slime plants have adopted filtration in place of decantation, and obtain extractions ranging from 92 to 95 per cent. From the above Rand figures, about 4 cents should be deducted for precipitation, refining and assaying.
Cohparattve Agitator Data
M. — Bottom Drive Mncbanical P.— Pachuca D-— Dotr
Sp. Grav. of PulpHl.32e Power at S6 per U.P. per moDth. Ltibor at S3 per «^t-hour shift.
Tona oJEUiDie.
Type.
•Xt
M
30X12
D— 3.0 Hp.
D— l.Scenta
D
30X12
M— 6.2Hj».
P— 2.3 oent*
P
15X45
P— 6.9 Hp.
M— 2.4oentA
a
/
Al.*rTABlLtTT TO
9a
s
ff
l|
¥
n
p
il
1
In
P
D
M
D
p
p
D
D
P
D
order of
D
P
T>
P
D
D
P
P
D
M
preter«.oe
M
M
P
M
M
M
M
M
M
P
The rich lead must be again cupelled, this time to remove impurities such aa copper and lead, using an air blast for oxidizing. A fine gold bully
lion Ib produced which is remelted in a Faber du Faur retort furnace (see Fig. 156), to yield ingote or bars of 380 fine in gold.
At the refinery of the Goldfield Cons. Co. they have adopted blast-furnace smelting of briquettes, a mixture of the precipitates 100 parts, Uthai^ 100 to 125 parts, and heavy sulphides concentrate from the mill (S 0.35 per cent; SiOa 30 per cent) together with some mill sweeping. The furnace yields a lead bullion (cupelled as above described, also copper matte and slag. The matte and slag may be sold to a regular smelting works.
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Chapter Xvi Typical Gold Mill Practice
The following pages give descriptions of various cyanide mills: CYANIDING FREE-HILLING POROUS ORES
These include ores which are somewhat porous and which may be cyanided by leaching when coarsely crushed.
THE WASP no. 2 MILL, SOUTH DAKOTA
The ore is a massive iron-detained quartz.
Its average gold content is $2.40 per ton. The ore is mined by steam-shovel, and crushed by two No. 6 and one No. 4 Gates crushers, to IJ-in. size. Four sets of 16X36-in. rolls reduce the ore to i-in., when it is elevated to storage bins. The rock is fed from the storage bins through rack and pinion gates to an IS-in. rubber belt conveyor, mounted on a frame which moves back and forth on an 18-in. track, so that it can dump to any one of the six individual conveyors, each of which serves one vat, A 1 H.P. motor and rope drive serves the m^n conveyor, and a 40 H.P. motor the individual conveyors and feeders. There are six leaching vats 12x32 ft., holding 420 tons each, fitted with the usual filter bottoms. Cyanide solution is pumped to the vats so that it flows under the filters and up through the ore. Two or 3 lb. of lime is added per ton of ore for protective alkalinity. The solution contains 5 lb. of cyanide per ton. This stands for twelve hours, is drawn off and followed by seven weak-solution washes. Gold is precipitated by zinc shaving. Cyanide consumption averages about 0.4 lb. per ton, and costs are 67 cents with an extraction of 76 per cent.
Ores Of Clayey Nature By Ctahidino
Any ore which contains a high percentage of alumina is difficult to treat, and two good examples of this ore are the Buckhom, Nevada, and ^ctoriouB, Western Australia.
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Typical Gold Mill Practice
The Victorious Hill, Western Adstralu
The lodes of the Victorious mine are of soft kaoUnized material, through
which run small veins of ironstone quartz, which carry the gold. This ore
has been successfully treated by the following process, of which a flow
output averages 320 tons.
Coarse Crushing.— The ore is broken by a 10X18-in. Blake-type rock-crusher running 250 R.P.M. After being crushed the broken rock passes to a 14-in. Robins belt conveyor, 'ikaSiMoa ^* *^ *" angle of 20° and is distributed into "^ '■* a bin by means of a Robins 14-in. hand-tripper with a double chute. The total storage capacity is 600 long tons. The bin 18 fitted with rack and pinion doors and steel chutes.
The ore is ground by four Wt, Huntington mills fed by four ore-feeders, driven by means of a short belt from the Huntington mill shaft. Cyanide circulating solution is used for crushing, and mercury is used in the mills with the present ore. The product of the mills equals 1 ton of ore to 1 ton of solution, and is discharged into a cement launder and thence to the pump well. From here it is lifted 35 ft, to the top of four sets of cone separators by a duplex plunger pump. The sand separated is conveyed by launders to a pair of Wheeler Flo. 120.— Flow-flheet, Victorious P^ns- Here more coarse gold is collected and Mill, Western Aiistralia, the sand is ground. The overflow from the grinding pans joins the main body of pulp from the separators in a collecting-box situated 4 ft.' above the thickeners, from which it is distributed to four pulp thickeners. These are steel vats 25 ft. diameter by 9 ft. deep. The agitators are 20 ft. diameter and 6 ft. deep and the arms revolve at 7 R.P.M, From the agitators the pulp is pumped to a distributing agitator near the filters. Filtration is done by three Ridgway filters whose capacity is estimated at 300 tons per twenty-four hours on this ore. This machine has a large shaft carrying two arms on which are suspended a basket of filter-leaves. This is dipped into a pulp tank to form a cjvke, then turned into a wash tank, and finally discharged into a hopper in the center. The ore averages $7 per ton, and an extraction of 90 per cent is obtained at a total cost of $1.10 per ton.
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The City Deep Mill
The Eolar Field
The ore mmed at the Kolar field, India, is free milling. The percentage of pyritic content varies on the different mines, but taken as a whole it does not average more than 1.5 per cent and for this leason the sand can be weathered without detriment. The weathering, while oxidizing the pyrite, and freeing the gold content, is further advuitageous in freeing the sand of surpluB moisture reducing it frcon 12 to 16 per cent down to about 3 per cent. This converts it to a more friable state, in which any shme present can be easily powdered. Having no water to displace, it can thus be easily saturated with solution, and the lixrviation and subsequent washini^ more thoroughly and quickly carried out.
THE CITY DEEP HILL This treats ore from the reefs on the Rand which may be described as a conglomerate of quartz pebbles cemented together by a matrix having a dark bluish appearance when freshly mined. The "banket" carries up to 75 per cent sihca and 2,5 per cent pyrite. The ore from the mine is divided into two classes by screening through grizzlies; the fine is conveyed by a 20-in. belt direct to the main ore bin. The coarse is taken by four inclined sorting belte, see Fig. 121, where the waste rock is picked out by-hand, each belt feeding three crushers with 12x24-in. jaw-opening. The return portion of each sorting belt receives the rejected waste rock and delivers it to a belt that takes it to waste. The transport of ore to the mill and all necessary surface work is effected by heavy electrical locomo- ^"- >2I.-S..rting Belt.
tivea using 2000-volt, 50-cycIe, 3-phase current. The mill of 2200 tons daily capacity is equipped with 200 stamps arranged in units of ten, each unit driven by a SO.H.P. motor. The weight of the stamps, which have long heads and short stems, is 2000 lb. One special feature is that there is only a layer of ^-in. felt between the mortar bases and concrete foundations. King-posts are entirely dispensed with, the concrete foundations being carried, with indented steel-bar reinforcing, to above the level of the top of the mortar-box. Each cam-shaft is carried by a steel frame and rests on eleven bearings, so as to minimize the risk of breaking. Stems are 4 in. by 13 ft. long and the stamps are arranged for a heavy duty if n
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192 Typical Gold Mill Practice
Each battery is provided with four Chalienge feeders. Referring to diagram, Fig. 122, the watery pulp, after elevation by a sand pump, is classified in Caldecott diaphragm cones, Fig. 58, and the underflow delivered to the tube mills. There are nine of these, 5) ft. diameter by 22 ft. long driven by a 100 H.P. motor. The tube mill dischai^ is conducted to amalgamating tables in the gold-reeoveiy house and under the same roof are arranged the extractor or zinc boxes, the clean-up machinery, strong room and the refinery so that all operations are performed in one building under the supervision of a responsible man. The pulp from the table is a^n elevated to four coarse sand classifiers, the coarse underflow of which goes back to the diaphragm cone for regrinding. Meanwhile the overflow of three classifiers passes to the slime separators, the overflow to slime collectors as described, the sand underflow to the sand collectors consisting of six steel vats, 50 ft. diameter by 10 ft. deep, set on reinforced concrete supports. A 24-in. conveying belt, running
FiQ. 122. — Grinding and CliiBH[fying on the Rand,
under the center line of these vats, takes the sand excavated from them by a Blaisdell excavator and by conveying belts removes it to twelve leaching vats. The sand is fed to these leaching vata by a Blaisdell distributor, and, after leaching, it is discharged by a Blaisdell excavator through a central discharge opening at the bottom of the vats to two 24-in. conveying belts, one under each row. A cross belt delivers it to a 24-in. inehned belt which takes it to the tailings dump, where it discharges 100 ft. above the ground.
The slime is collected in four conical-bottom steel v&t^ also on reinforced steel supports. From there it is taken for treatment in two steel conical-bottomed air agitator vats, 32 ft. diameter by 38 ft. deep and later washed in eight steel conical-bottomed vats 70 ft. diameter by 16j to 20i ft. deep. This washing is done by decantation.
It is thus seen that from start to finish the labor of the South African native has been eliminated as far as possible. The various products either flow from point to point by gravity or the fluids are pumped and the soUd products are transferred by mechanical methods.
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Consoudated Langlaaqhte Mill
THE HILL OF THE CONSOLIDATED LAnCLAAGHTE CO^ RAND, SOUTH AFRICA
The mill is designed for a capacity of 45,000 tons per month of twenty-ax days. The receiving bin, which ia of steel and concrete construction, delivers the ore on to three 30-in. belts running at a speed of 150 ft. per minute. E^h belt discharges over a short grizzly into a washing trommel. The imdersiEe from the grizzhes and trommels discharge on to a 30-in. belt, which dehvers on to the main 30-in. conveyor belt leading into the mill. The washings from the trommels and the sorting belts are conveyed by launder to the coarse sand pumps. The oversize from each of the trommels is delivered to a 36-in. sorting belt, running at a speed of 40 ft. per minute. Waste rock and tube^nill pebbles are thrown into separate bins under the sorting belts. Each sorting belt discharge to jaw crusher, the product from which is carried by a 24-in. belt to the mill conveyor belt. The dust produced in breaking the ore is exhausted by a fan, and, after spraying with water ia delivered to the launder carrying the tronmiel under-size. The sorted waste is transported from the bins by trucks to the dimip. The ore is elevated at an angle of 18° by a 30-in. belt conveyor, running at a speed of 300 ft. per minute, and ia distributed over the 3000-ton mill bin by means of a tripper. The tube-miU pebbles are conveyed in a umilar manner from the crusher station and thence to the pebble-storage bin at the west end of the mUl building. The mill, is supplied with 100 stamps each of 1750 lb. Ten stamps are driven by a 50-H.P. motor operating through a counter shaft to two cam-shafts, each of which carries five stamps.
There are ten fiXie^-ft. tube-mills, each driven through spur gearing and belt by a 100-H.P. motor. The battery pulp from each ten stamps gravitates to a 6-ft. primary cone classifier at the head of each tube-mill. One secondary cone of smaller size is provided for each two primary cones. The imderfiow from the primary cones passes to the tube-mills, and the overflow passes into the secondary cones. The overflow from the latter is led direct to the fine-sand pumps, while the underflow is led to the coarse-sand pumpe. The tube-mill pebbles are brought from the pebble-bin by cars and are deposited into a hopper placed at the inlet of each tube-mill. In the platehouse there are forty stationary amalgamating tables, each 5X7 ft., and recovery by amalgamation ia 72 per cent. E^h tube-milt discbarges through a launder on to a set of four tables. Opening out of the platehouse is the clean-up room, which ia equipped with two retorts, two bullion furnaces, three amalgam barrels, batea, and amalgam press. A strong-room is (Jso provided. The pulp from the plates is conveyed by a launder to a reinforced concrete sump having a capacity equal to the product of ten minutes* crushing in the mill. The pulp is elevated by 10-in. pumpe with 12-in. suction and is distributed to the tube-mill primary
194 Typical Gold Mill Practice
cones by means of launders. The overflow from the secondary cones paesee as mentioned above, direct to the fine-sand pumps. These pumps are provided with suction hoppers, having overflow launders leading back to the coarse-pump sump. The fine-sand pumps elevate the final pulp to the sand classifiers. The principle of double classification is adopted, the classifiers being cones 8 ft. diameter by 10 ft. deep. The primary cones, six in number, are fitted with rings to catch candle grease and wood fiber from the mine. The underflow, suitably diluted, passes to three secondary cones, which are fitted with water regulators giving a wet underflow. The underflow from the secondary cones passes to three sand-collecting tanks, which are each 55 ft. diameter by 10 ft. deep, and are provided with peripheral launders. Butters distributers are used. The overflow from the primary classifiers is led to four sand retxim cones, each 7} ft. diameter by 5 ft. 10^ in. deep. The percentages of sand and slime are 39.5 and 60.5
Fio. 123.— St«el Tank, 70 ft. diam., for Rand Practice.
respectively. The overflow from these cones, together with the overflow from the secondary classifiers and the sand collectors, goes direct to the slime collectors and the underflow is returned to the fine sand-pumps for reclassification.
The sand is dischai^^ed from the collectors by bottom discharge doors on to belt conveyors, which carry it to a Blaisdell distributer, placed over the treatment vats. There are twelve treatment vats, each 40 by 9} ft., and the treated sand is discharged into cars and conveyed to the dump by mech^cal haulage. Recovery by sand treatment is 12 per cent. The extractor bouse contains zinc boxes and is provided with an acid-treatment plant and smelting-room. The tatter contains calcining and reverberatory furnaces, ball-mill, and strong-room. The decantation sUme plant consists of the following tanks: two collecting tanks 70 ft. diameter by 12 ft. with 7.ft. 6J-m. cone, see Fig. 123; five collecting tanks 50 ft. diameter by 12 ft. with 6-ft. cone; two treatment vats 70 ft. diameter by 12 ft-with 6-ft. cone; one intermediate transfer tank 50 ft. diameter by 12 ft. with 7-ft. fij-in. cone; ten treatment vats 50 ft. diameter by 12 ft. with
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The Homestake Mill 195
6-ft. cooe. Recovery by slime treatment is 12 and total extraction 96 per cent. The overflow water from the collectors is led to a return water sump, from which it ia pumped to the mill supply tanks by 10-in. pumps. The residue is worth 23 cents per ton.
The Houestake Hill, Lead, South Dakota
The ore is a gametiferous hornblende schist cwrying 7 to 8 per cent of [^rite and pyrrhotite blether and of the value of $4 per ton. It is treated 1^ amaigamaiitm and qfaniding. The coarsely crushed ore is fed to 640 stamps having, as shown in the flow sheet, Fig. 125, an output of nearly 2800 tons per day. Here it is crushed in cyanide solution in the ratio of eleven parts solution to one of ore, A tube-mill is here introduced to increase the fineness of the stamp-battery product. The battery pulp passes over in all two-thirds of an acre of amalgamating plates set upon a 12^ per cent slope of 1^ in. per foot. With inside amalgamation in addition to these plates 72 per cent of the gold is caught. The tailings now flow to fourteen gravity cone classifiers /, each 4 ft. diameter by 5 ft. 5 in., which yield an overflow going to the aand dassijicalion ayatem and an underflow of 8 per cent of the ore which goes to the regrinding system to be r^round in tube-^nills. All the pulp is of such a nature that a solution contact of some four to eight hours ia ample. Direct filter slime treatment is, perhaps, the most efficient and satisfactory practice of the day. The partially thickened slime, in water, is charged directly into the filters, where the slime cake is in an excellent condition to receive preliminary treatment, such as aSration, solution leaching, and washing. The entire dissolution ia efifected in the filter, with a minimum amount of solution, precipitation, consumption of chemicals, power and labor, and loss in dissolved gold. The plant is very simple and compact, while the various operations are susceptible of accurate technical control. The pulp from the grinding machines unites with the main stream of pulp before arriving at the second battery of classifying cones. Classification of crushed products into sand and slime is effected by means of four series of sheet iron cones with 50 to 80° slope with peripheral overflow, each unit discharging at the apex through a short cast bushing. Nearly all of the material overflowing from the various cones of the classification syt^tem will pass a 200-mesh screen, and is treated as slime. It is thickened, and then run to the slime plant storage tanks, lime is added and treatment continued as described later.
The Sand-plant. — The prepared sand contains 40.5 per cent coarse particles that remain on a 100-mesh screen ; 30.8 per cent middling, between 100 and 200 mesh, and 28.7 per cent fine passing 200-me8h. This leaches at the rate of 3 or 4 in. per hour. Before the sand enters the leaching vats it receives a stream of milk-of4ime which has been prepared by being
Typical Gold Mill Practice
stamped in a one-stamp battery reserved for the purpose. From 4 to 5 lb. of the lime is added per ton of sand. The classified pulp and lime thus mUed pass to a Butters distributor, which can be transferred from one vat to another by an overhead trolley- There are 20 leaching
vate, each 44 ft, diameter, 9 ft. deep, and capable of holding 600 tons. The vat is filled with water and the sand runs in. It takes nine hours to chai^ the vat, and treatment lasts five days. When the vat is filled the ore is drained and a series of washes of the stronger of the stock soluiaons
The H0Me8Take Mill 197
(containing 0.14 per cent KCN) is run in, allowing each wash to dmin off below the top of the ore to draw in air. Besides this, air is introduced below the filter. The effluent, its strength reduced to 0.10 per cent, is run to the two weak-solution precipitation tanks /,/, Fig. 125, each 26 ft. diameter by 19 ft, deep. After this, the weak solution is brought upon the charge and rettuned two days more. The solution escaping during this period is run to the two strong-solution collecting tanks, e,e. This is followed by a water wash which finally reduces the unextracted gold to 5 to 7 cents per ton.
The charge is now ready for shucing out. This is done by two men in 2i to 3i hours, four side gates and one bottom gate being used for the purpose. The 8-oz, duck filter-cloth underlaid with another of cocoa matting h washed clean. The vat is then filled with water and is ready for t^e next charging.
The Slime-plant — The slime pulp, amounting to 1600 tons daily, has an average value of 91 cents per ton. It contains 3 tons of water to 1 ton of solid, and is carried two miles by a 12-in. pipe at a grade of 1 .5 per cent to the slime-plant. Here, two small vats are provided for slaking lime. The content is drawn to a screen-bottom box where the xmdissolved lumps separate. The box overflows into an agitator from which the milk-of-lime continuously runs into the main slime-stream at the rate of 5 lb. of lime per ton of dry slime. Two storage vats, 26 ft. diameter and 24 ft. deep, having conical bottoms with 47° aides, receive the stream. From the bottom of these storage vats the slime-pulp is drawn continuously through a 10-in. pipe to large Merrill filter presses, 65 ft. below, to obtain a pressure of 30 lb. per square inch. The U-in. main extends the whole length of the press-building. Between each pair of presses the main branches into 10-in. pipes, which in turn send two 4-in. branches to each press. The smaller branches connect to a 4-in. passage or channel that extends along the center of the top of the filter-press frames. From the channel the slime-pulp flows into the press. There are ninety-two frames each 4 by 6 ft. and 4-in. distance-frames to form slime-cakes 4 in. thick.
Collecting is by revolving distributors, while the residues are sluiced out with 1 ton of water per ton of sand.
Total cycle approximates seven days. Screen sizing of sand: 35-mesh (0.0164 in.), 23 per cent; 35-65 (0.0082 bi.), 29 per cent; 65-100, 21 per cent; 100-200, 18 per cent; 200, 9 per cent.
Extraction on S3. 15 sand heads, 79.9 per cent.
Coat of sand treatment per ton of sand, exclusive of neutralizing, precipitating, refining and assaying, 23.5 cents.
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Typical Gold Mill Practice
COST OF SARD TRBATHBNT PER TOn OF SAND
Rand
C«ili
Ubot
AMftying, aftmpling 0.33
lime 0 56
Cyanide 7 . 23
Precipitation ■. 1,56
Zinc ahavingB 1 , 2S
Clean-up smelting 2.92
Sodium bifulphate O.M
Miecdlaneoua supplies 2 . 30
Water 1,60
Misoellaneoui 0.59
Total 40. M
Homrtlake
Cent!
Superintendence 1 , 06
Assaying 48
Neutralization 1.54
Transportation 0. 17
CItksaification 1 . 18
Treatment 8 G9
Precipitation and pumping solutbns 1.18
Re6niiig 38
Heating 0.51
Miscellaneous 0.81
Repairs 1.88
Total 17.78
TOTAL COST OF HOHBSTAXB DIRECT-FILTBR SLUIE TREATMENT PER TON OF SLIME
Superintendence 0.93
Assaymg 31
Thickening 34
Treatment 10.73
Piecipitating and pumping solution. 1.66
Heating 17
Miscellaneous 88
Total 18.38
Or, combined as:
Cota.
Works labor 6.70
Shop labor 84
Power 1.84
Chemicals 6,78
Total 18.38
Consumption Per Ton Of Slimb
Um.
Sodium cyanide 0. 161
Zinc dust O.lia
lime 3.840
Ha 0.S93
Power in Kw. Hrs 1 . ISO
Actual cost of dissolulion, deducting precipitation, refining, assaying, and one-third of superintendence and miscellaneous items 0.1516
Capital Cost — AsBtiming an average total cycle of six hours, the complete capital cost of this plant including classification, thickening, filtere, precipitation, refining, pumping, heating — all enclosed in buildings — will approximate $265 per ton of daily sUme capacity.
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The Liberty Bell Mill
THB LIBBILTT BELL Hnx, TBLLDIUDB, COLO.
The mine i»oduce8 a hard ore containing 85 per cent alica, 10 per cent lime and about 4 per cent pyrite. The treatment consists in amalgamation, concentration and cyanide treatment of the concentrates.
The ore coane-cnished at the mine to 3-in. size is delivered by tramway into the battery ore bins, see the plan, F^. 126. The mill contains eighty stamps of 850 lb. each in eight batteries set on concrete foundations. The ore, fed from the bins by Challenge feeders, is wet stamped to 10-to 12-mesh in a solution containing 2 lb. cyanide per ton. For each battery there are two plates 4 by 8 ft. set across the flow, the first plate with a grade of 2) in. to the foot the second Ij aq. in. The slower flow on the flatter plate ensures recovery of the finer gold particles. The pulp is now treated in four Richards three^pigot vortex classifiere followed by six 6-ft. settling cones, the spigot discharge being concentrated on 18 Wilfley tables. The concentrate is reserved for special treatment by fine grinding and cyaniding while the tailii^ pass on to be reground in three tube-mills 5 ft. by 22 ft. long. The discharge of the tube^nills is re-amaJ^^amazed by being flowed over ei^t sets of transversely set amalgamating plates where the remaining gold particles are caught, and reconcentrated on ten Deister slime tables. The concentrate is united to the first above mentioned and' the slime tailings are combined with the overflow of the cones into one pulp to go to the nine 33 ft. diameter hy 10 ft. deep Dorr thickeners, where adequate settling of the slime can be effected. The tUckened underflow or Bp^ot dischai^ of these goes through six Hendryx agitators, 17 ft. diameter, operated continuously in series, the discharge of the last one being pumped to the equalizer tank of the filter system. Here it is drawn off as needed into any one of the five Moore suction filters resembling the Butters, The flow or strong solution from the Moore filters passes to solution storage thence to the sixth filter used as a clarifyit^ filter and so on to the precipitating house where the zinc boxes are. The weaker solution resulting from the washing of the filters also goes to the clarifying filter and to other dnc boxes, the discharge being pumped to a tank to be used in the Moore filter for blowing off the cake. The precipitate from the sine boxes is acid treated in the precipitation house and with the retorted amalgam of the plates is melted, in the melt house and shipped to the Denver mint.
Tseatmbnt Op Tblldkide Orbs
This class t^ ore is mined at Cripple Creek and Kalgoorlie, but the percent^^ of telliuium is now quite small, the bonanzas having been worked out for several years.
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Typical Gold Mill Practice
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1
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1
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The Golden Cycle Mill 201
Thb Golden Cycle Mill, Colorado Springs, Colo.
This is a custom mill treating Cripple Creek telluride ores, containing
no other deleterious metals, by roastiog and cysniding. The ores,
varying in size from 1.5 in. diameter to &ne sand, after sampling, are
takeo by belt conveyors to one of the three lai^ bins, called bed-
Fiu. 127— Flow-sheet ot Golden Cycle Mill.
ding floors. They are classified into the grades A and B, according to the lime content, and separately and uniformly distributed in large beds of 5000 tons, so as to enmire a good miidiiuc. Class A contains Si02, 76.7 per cent and CaO 1.57 per cent; class B all over that, a typical analysis of the latter being, insolublematter 75.9 per cent; AI2O3 3,4 per cent; Fe 3.5 per cent; CaO 5.1 per cent; S 1.8 per cent; MgO 1.1 per cent. Class A resembles it except in the lime content. From the bedding floors the ore passes by belt conveyor to six ball-mills and is then dry crushed to pass a )>in. screen and with an average of not more th(Ut 30 per
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202 Typical Gold Mill Practice
cent coarser than lOnoieah. The basic ore8 need finer crushing than the silicious ones. The united capacity of these mills in 1250 tons daily. The ore from the mills is carried by a belt conveyor to steel bins set above nine duplex Edwards roasters, Fig- 74. This furnace 165 ft. by 13 ft. wide has a roasting hearth area of 1495 sq. ft. The rabbles revolve 6 R.P.M. There are three fireboxes to each roaster. There is a cooling hearth 44 ft. long by 14 ft. wide or with an area of 672 sq. ft. In roasting class A ore, the temperature in the flow of heat at No. 2 firebox is 800° to 850° C, For the claes B ore, a higher temperature of 850° to 900° C. is maintained. The ore escapes from the roasting hearth at 485° C. and leaves the cooling hearth at 278° C. For roasting class A ore 12.5 per cent of fuel is used, class B, containing so much more lime, takes a higher heat and at least 17 per cent of a local lignite coal. E^h furnace roasts 125 to 150 tons of class A ore, 80 to 100 tons of class B ore daily. The roasted ore falls from the cooling hearth upon a reciprocating drag conveyor, where it is sprayed with water to cool it to about 90" C. (194° F.), so that it may fall upon the rubber conveying belt to be taken to bins set above seven 6-ft. Chilian mills where the cyanide treatment begins. Here the ore is fed with addition of cyanide solution to the mills and ground to pass a 50-raesh screen. The discharging pulp is distributed upon the tables, blanket-covered and Bet at a slight slope, where the pulp spreads out in an even layer and where, with the aid of some solution, the lighter pulp is washed away, while the coarse gold developed in the roasting is cai^ht in the interstices of the blanket. The pulp now passes to the bowl-type Dorr classifiers where it is separated into two products, clean sand for leaching, and slime, which is agitated and the gold solution removed by vacuum filters. One may note here the beneficial effect of roftsting for not only is the telluride of gold decomposed, but the colloids of fiocmilcnt portions are shriveled up by the heat. Precipitation is effected by sine shaving in the usual manner. Cripple Creek ore, if slimed and given a long period of agitation, will yield aa much as by the above-described method. Yet roasting prevails, since, while it costs 50 to 60 cents per ton, it is not necessary to grind so fine, the extraction is a matter of hours rather than of days, cyanide consumption is leas, and the solutions are less likely to become foul.
THE VICTOR PLAUT OF THE PORTLARD GOLD MIITmO CO., VICTOR. COLO.,
was built for the treatment of the ore from the Portland mine which would not withstand the high cost of freight and treatment when shipped to the Portland plant at Colorado Springs.
The ore is brought to the mill in 5-ton electric cars and dumped into a cylindrical steel bin above the crushing plant. From this bin it is fed by an apron conveyor to a 15 by 30-in. Blake crusher, which reduces the ore
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Thb Victor Mill 203
to about S-in. size. It then passes to a 36-iii., style B Symona disk cruBher, which machine reduces it to Ij in.; thence to a set of 20 by 48-in. rolls, the entire product of which will pass a 1-in. ring. A belt conveyor takes this 1-in. product to the main mill building, where, after passing throu^ a Veein sampler, it is distributed into four steel storage bins. These four bins discharge by plunger feeders to four Wt. Akron Chilean mills. At this point a weak cyanide solution is introduced, the mills discharging a pulp through a 30-mesh screen, which flows and is distributed to thirty-six Wilfley tables. The concentrate from these tables is finished on six Wilfley finishing tables. This last set of tables yield heads, high in iron sulphide and c<»>taining some gold which goes to the smelter, and a siticious tailings, which after sliming in a tube-miU is mixed with the regular mill slime.
The tailing from the Wilfleys runs to four Aldns classifierB, where it is divided into sand and slime. The sand goes to a continuous wash system (Akin classifiers) whence, after being washed free of soluble gold, it is hauled to the dump. The slime is pumped to thickening cones, where, after thickening it is reconcentrated on Card tables. The concentrate from the Cards join that from the Wilfleys. The tailing from the Cards runs to the Aldns thickeners, the thick pulp from the same going to air agitators and thence to Portland filters, whence, after being washed free of soluble gold, it is hauled to the dump. The e£9uent sc^ution from the Portland filters joins the clear overflow from the thickeners is clarified, treated by the Crowe vacuum method and by the Merrill precipitating process, and goes to the zinc-dust precipitating plant. The mill has a capacity of 500 tons daily, and uses only 1,000,000 gal. of precipitated water per month.
Kalgoorus District, Vbstbrh Australu
The ores ate silicious and contain pyrite and tellurides of gold and silver. There is but little &ee gold and that mainly in the pyrite. At the Ka^oor-lie mills two systems of treatment grew up, the dry or roasting process as described for the Portland plant at Colorado Springs and the wet process. It soon was found that the telluride ores were not soluble in ordinary cyanide solutitms, but bromo cyanide proved to be effective (see " The Bn»no-cyanide Process.")
Grinding pans appear to hold their own in the treatment of telluride ores in the Kalgoorlie district. They are used for intermediate grindii^, but for sliming, tubennills are considered preferable to pans. The For-wood-Down pan has a classifying discharge, the pulp issuing from the pui by a row of 1-in. holes near the top. Outside the piui is a pocket or launder having a slot at the bottom leading back to the pan. Heavy sand
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204 Typical Gold Mill Practice
settles in this pocket and is returned through the slot while the fine flows over the outer edge of the pocket.
Gradually the filter-press on the Kalgoorlie field is being discarded in favor of vacuum-filters. In the past an inunense sum has been spent on press plants, about 100 presses being erected, 75 of which are treating 100,- 000 tons monthly. The others are out of commission, but the benefit derived from this machine at Kalgoorfie has been admittedly lai^. With a press, washing can be carried to a degree that cannot be beaten, but the labor cost is high, and it is expected that the press will have to give way. The Associated Northern, Boulder, and Oroya Links filter their current mill slime by vacuum systems of their own, while others are talking about introducing the system.
The Orota-Brownhill Mill, Kalgoorlie District
This uses a wet process of concentration followed by cyaniding. The ore after coarse crushing in a rock-breaker goes to the stamps. Here it is crushed in a weak solution of about 0.4 per cent or O.S lb. per ton of cyanide, kept alkaline by the addition of lime and the pulp r\m to hydraulic classifiers. Here in closed circuit with Wheeler pans. Fig. 135, the sftnd is finely ground, the classifier overflow then going to Wilfiey tables. The concentrates from the tables receive a special treatment as follows:
These amount to 6 per cent of the ore milled and contain 11 ounces gold per ton and the great bulk of the refractory elements in the ore, thus leaving a tailings product well suited to subsequent cyanidrag. The concentrate is sent to three single-deck Merton roasting furnaces similar to the Edwards furnace (Fig. 74), where it is roasted with from 0,5 to 2.2 lb. salt per ton, each furnace easily roasting 10 tons per day. The roasted product is removed by push conveyor aa shown in the Edwards roaster to two pairs of Forwood-rDowna 5-ft. pans in parallel. The' first pair is used both for fine grinding and amalgamation, their overflow passing to the second pair used for fine grinding only. Mercury is added to the first pans three times daily, amalgamation recovery is about 30 per cent of the value of the concentrate. The second set of pans give an overflow which passes on for cyanide treatment. It is agitated in vats with mechanical i^tators with a solution of 0.1 per cent or 2 lb. per ton for 100 hours, lime and lead acetate being sometimes added in the pans. The pulp is afterwards filter-pressed. The pregnant solutions are precipitated in ordinary zinc boxes using zinc shavings, that have been dipped in lead acetate solution.
Returning now to the tailings from the Wilfley tables, these are classified in closed circuit with tuiie-milla, so as to give an all-slimed product for cyaniding. This is dewatered or thickened and goes to the agitation
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The Holunger Mill 206
vats having mechanical stirrers. Here it is agitated with cyanide eolution of 1 per cent or 2 lb, cyanide per ton for three hours, the cyanogen bromide is added at the rate of about 1 lb. for each ounce of gold and the agitation continued for twelve hours. About two hours before the completion of the agitation quick lime, 2 or 3 lb. per ton is put in. After agitation the putp is discharged U> a stirrer or agitator to keep it properly mixed while it is being pumped info the filter-presses. The gold-bearing solution from the filter-presses passes through clarifying presses and thence to the zinc boxes for precipitation. All the tailings are taken from the presses to the dump by conveying belt. In 1905 when the ore carried $32 per ton in gold the extraction was 94.6 to 95 per cent.
The Hollinger Hill, Porcoputb District, Ohtahio, Canada
This is one of the new mills of the district and treats 499 tons of $20 ore dfuly, recovering 93 per cent of the gold. The ore is of quartz and schist with a high percentage of [)yrite. It is soft and easily crushed, yielding a heavy pulp which ^ves rise to some mechanical difficulties in the agitation. The gold is free. Treatoient consists in concentration, cyaniding the concentrates and die tailings. Fig. 128 is a flow-sheet of this mill. The ore is coarsely crushed in two stages, the product from the first or gyratory breaker being screened by a trommel having 2^-in. holes. The oversize of this is crushed by a 20-by 10-in. Blake crusher so that a product of less than 2^ in. passes by belt conveyor to be distributed to the mill bins. Tbence it is drawn off in regulated quantity by Challenge feeders to the stamp. It will be noted that the feed bins hold 2^ days' supply of ore. Stamp crushing is done in a solution of 1.5 lb. cyanide per ton, umng 5 tons per ton of ore crushed. The ore is now crushed to 6-mesh size through Dorr classifiera taid tube-mills in closed circuit. The overflow from the classifier is first again treated by means of large Spitzkasten, 20) ft. long, 6 ft. wide, and 6 ft. deep to yield two products, an overflow for direct cyaniding and an underflow, a product that undergoes extensive concentration in order to recover the free gold. The imderfiow of 25 per cent solids is concentrated upon Deister slime tables similar to the Wilfley table (Fig. 84) yielding concentrates largely pyrite, containing the free gold and a tailings which joins the Spitzkasten overflow for cyanide treatment. The concentrate, comparatively free from solution, is taken by a spiral screw conveyor and bucket elevator to four small bins ready to be fed into Wheeler pans in charges of 1) tons of the concentrate with 100 lbs. of mercury. After grinding for several hours the contents of the pans are discharged to 8-ft. settlers, Fig. 136, where the tailings overflow for cyaniding while the amalgam collects in the mercury well of the settler. Referring now to the various overflow products for cyaniding, these are pumped to Dorr
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Typical Gold Mill Practice
Fia. 128. — Fkiw-eheet of HoUinger Mill.
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The Tom Reed Mill 207
thickeners which deliver an underflow containing 50 per cent solids. With an addition of barren cyanide solution this product is agitated for forty-eight hours by Trent agitators in series and is then filtered by Moore suction filters to give a tailings that is rejected still holding 25 per cent moiHture. The clear overflow of the Dorr thickener is joined by the filtrate from the Moore filters, but must be clarified before precipitatitm. Precipitation, using sine dust, is performed as described under the head of " Merrill Precipitation Process." The filtrate returns as barren solution to the battery storage tank near the stamps.
The Tom Reed Hill, Oathar, Ariz.
The treatment is by cyaniding using continuous counter-cuxrent decantation. The gcdd occurs principally as hematite in quarts ore of low grade, but in large bodies. Referring to the flow-sheet, Fig. 120, the run of mine ore crushed through a gyratory and a Dodge jaw-crusher in series is reduced to a maximiun use of 2^ in. and is taken t^ an inclined belt conveyor to the feed bins. A 16 by 16-in. steel chute at the bottom of the bin makes it pbssible to draw off the ore to a Stephens-AdanwHi apron feeder to a 6 ft. by 5 ft. ball-mill which grinds in closed circuit with a Dorr classifier, with addition to eatae pregnant cyanide solution from the decantation tanks. The overflow from the classifier passes to two pairs of similar classifies on the floor below, each pur in closed circuit with a 5 by 6-ft. ball-^nill. Here a further addition of pregnant solution results in a product S5 per cent of which is less than 200-mesh size, this fine grinding being necessary to ensure contact of the gold by the cyanide solution. The pulp thickened in the primary 40-ft. Dorr thickener to a specific gravity of 1.5 is raised to the first of four 40-ft. agitators, where it is agitated m series with addition of air. From the last of these machines the pulp passes to the lower of two Eronier sand pumps. The lower pump delivers to the upper one and that in turn to a distributing box. From this box, the pulp is divided between the head tanks of two series each of four Dorr thickeners, Rg. 103, operating on the continuous counter-current decantation system (C. C. D.). From the last of the aeries the residual taiUi^ go to a settling pcnd. The overflow trom Uie head tanks flows as already mentioned to the ball-miUs and through them to the primary thickener. The nearly clear overflow of this thickener passes on to a vacuum classifier filter and after treatment by the Crowe vacuum method is precipitated with rinc dust by the Merrill precipitating process in a building near the barren solution sump. The filtrate from the Merrill process flows t^ gravity to two 6 by 6-ft. measuring tanks that are alternately filled and emptied by the action of a tilting laimder operated by two floats, one
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Typical Gold Mill Practice
in each tank. This launder in turn opens the discharge of a full tank and closes the empty one whOe at the same time it diverts the flow to the latter. The number of tanksful is shown by an automatic counter. From these
Pio. IW.— Flow-eheet of Tom lleed Mill.
tanks the solution goes to the barren solution tank 27 ft. diameter by 5 ft. deep. The barren solution is piped to the third tank of each dccantation series. Wash-water, in quantity sufficient to replace that discharged in
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The United Eastern Mill 209
the taOingB, is added in the fourth tank of each series. For refining there are provided two miiffles where the precipitate is dried and roasted in pans and a Steel Harvey tilting furnace Fig. 3, where the roasted precipitate is melted down into an ingot.
THE UmTED EASTERH HIU,, OATHAN, AKIZ.; SOO TONS DAILY CAPACnT
The ore is a mixture of calcite and quartz with some undecomposed andesite. It contfuns 1 oz. gold and but 0.34 oz, silver per ton. The gold is so finely disseminated that fine grinding is essential. Figs. 130 and 131.
^vatmenL — This consiets of (1) single-stage coarse crushing; (2) two-stage ball-milling in cyanide solution; (3) combined air and mechanical agitation; (4) stnught counter-current agitation; (5) removal of air by the Crowe vacuum treatment; (6) precipitation by nine dust using the Merrill process; (7) fluxing and melting the wet precipitate into bars.
With the exception of the Merrill filter-press for separating the precipitate no filters are used about the mill. It is remarkable that such a high-grade ore can be well extracted by counter-ciirrent decantation.
Coarse or Prelmunaiy CruBhmg. — All hips of ore that will pass a 10-in. grizzly are sent to a gyratory crusher, chosen rather for a sufScientsided jaw-opening for the grizzly discharge, than for its capacity in excess, this being 35 tons per hour. Lately the fines of this run of mine ore have been screened out by grizzlies set at 1 j in. opening, thus relieving the crusher and reducing crusher repairs. It is now thought that this crushing might better be performed in two stages for the production of a finer product ; it would secure a finer feed for the ball-milts that come next, thus increasing (A«V capacity.
One should note here that for an assured supply, storage for several days should be provided. This point is often overlooked.
Coarse Grinding. — From the feed bin the ore is fed to the ball-4nill by the traveling feeder, Fig. 320. The setting of a feed gate and the varying of the specific gravity of the pulp in a special Dorr classifier determine how thick a feed shall come out. The coarse g! jidtng is done in a Marcy ball-mill in closed circuit with a Dorr classifier (see Fig, 40), forged chrome iron balls being used in the mill. The grinding is done in a 1.6 per cent KCN solution, showing 1 lb. protective alkalinity.
Hoe Grinding. — This is done in two ball-mills 5 ft. diameter by 6 ft. long, also in closed circuit with Dorr classifiers, each mill being of 90 tons' capacity and giving a product, 82 per cent of which is minus 200-mesh. The mills both in the coarse and fine grinding discharge a product of 30 per cent moisture only.
Agitation. — From the fine^pinding department the pulp flows to No. 1 Dorr thickener, marked U on the flow-sheet. Fig. 131, arriving there witli
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210 Typical Gold Mill Practice
a Specific gravity of 1.12 and a dilution of 1 part ore to 4.5 parts solution. The thickened underflow from this thickener of a specific gravity 1.4 and amounting to 290 tons daily is pumped to the seven agitators Nos. 1 to 7
FlQ. 130.— Plan of United Eastern Mil).
of the plan and indicated by the single rectangle A of the flow-sheet, there being an addition of 197 lb. of NaCN to the first f^tator as tiiis is done. The period of agitation is sixty-two houra for complete solution.
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The United Eastern Hill
Aviation is carried on by ur at a pressure of 30 lb. per square inch, using two belt-driven compressors of 15 kw. each.
Thickening. — ^Fiton the thickeners A the pulp flows through five 40^. thickeners marked No. 2 to No. 6 on the flow-sheet and plan. These are arranged for straight couuter-current work. The pregnant solution from No. 1 agitator after passing the zinc-boxes becomes barren solution to go to No. 4 thickener and the wash water is introduced at No. 6 thickener. The flow at No. 6 is qilit and about one-third is sent to No. 7 thickener (see the plan). The underflow discharges of Nos. 6 and 7 Sow blether to the tftitiTigp pond at a moisture ctmlent of 0.82 ton solution per ton of ore.
No. 1 or primary thickener, which takes the dihite overflow fran the last Dorr claamfiers, has a settling area of 4.37 sq. ft. per ton of ore. It is necessary to operate it with a low mud line to prevent colloidal material getting over into the gold solution tanks. About 870 tons of solution pass to the taess solution vat while 600 tons go back to the mill storage.
Fio. 131.— Flow-sbeet of United Eastern Mill.
The refutation of the thickeners is maintained by varying the speed of the diaphragm pumps which feed them, the specific gravity of the dischfuge, and the depth ot the mud line being recorded every four houn. In maintaining alkalinity lime is added dry to the Marcy mill feed. The loss per ton is S3.215 and the loss of cyanide 0.625 lb. per ton of ore.
Cost (rf Plant— The cost of the mill proper is $133,539.09, to which should be added the crushing plant $11,975.37, the coarse ore bins $1,916.27, the refinery $5,789.13, and ih6 hme house $272.54, or a total of $153,492.40, which on a baeis of 200 tcms daily would be $767.46 or $2.13 per annual ton.
Operating costs are based upon $5 per day for helpers, $5.50 for mill men and $6 for sohition shift foremen.
In 1918 the cost of operations averaged fw operating labor $0.47, repair labor $0.09, supplies $1.06, power $0.52, miscellaneous $0.04, making a total of $2.17 per ton. Coaise crushing cost $0.07; coarse grinding $0.32, and fine grinding $0.50 per ton.
As respects particular costs also in 1918 per ton oi ore: Those for
212 Typical Gold Mill Practice
heating solutioQ were S0.03, of general expense 10.18, of lighting t0.008, of assaying S0.03, sampling S0.03, of tailing disposal S0.02, of clarification t0.04, and of precipitation S0.12.
The metallurgical report shows in 1918 that 92,339 tons were milled carrying 100,903.(Q of gold and 54,137.02 of silver, with a recovery of 96.75 oer cent.
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Chapter Xvii Treatment Of Gold Mill Concentrates
Classification
For this trealsaent it is rec(»nmended that for the grinding a tube-mill be used with a Dorr classifier in closed circuit. It yields an overflow, which passes on for agitation, and an underflow or spigot discharge, which goes back to the tube-mill. In this manner no granular particles can escape the grinding action of the tube-mill. The ground product goes to Dorr agitators followed by Dorr thickeners to be subjected to continuous counter-current agitation. Even then, before the tailings are sent to waste, they are Bltered and washed, using Oliver or similar filters. On the other hand, the clear overflowing pregnant solution of full strength is sent to the Merrill precifntation and clarifying process. The solution is handled by aid of triplex plunger-pumps. For the above methods of treatment a side-hill location may be used, with ample fall throughout with good dumping ground below for the tailings. This should be so arranged that the tailings can be stored, not carelessly allowed to run to waste. In the planning, the all-gr&vity arruigement is preferred so as to avoid the use of troublesome bucket elevators and pumping. The raising of solutions by pumping is another matter, even to raising them 60 or 70 ft. from sump to supply tank; there is no gnt in the sohition, a fruitful cause of piunp wear.
The concentrates which are to receive the special treatment above specified ccme to this special treatment plant, either in boxes filled from the head launder of the tables or in cars that have directly caught the product, or indeed into a launder that by aid of a stream of water carries it to the sump of a sand pump, this pump lifting it to collectii^ vats where it is dr^ed before subjecting to the special treatment.
The stuff can be treated in several ways.
(1) It may be treated with strong cyanide solutions in ordinary percolating vate for an extended time.
(2) It may be stacked to expose it to the weather, then percolating it as in method (1).
(3) It may be shipped at once to the smelting works.
(4) It may be treated by chlorination.
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214 Treatment Of Gold Mill Cwncentrates
(5) It may be all-elimed as outlined above, then agitated with cyanide solution and the tailings carefully filtered and washed.
(6) It may be roasted, then finely ground, agitated and the residue filtered.
A clean iron pyiite may be treated by method (1) with fair reeulte, but time, cyanide consumption, and the necessary retreatment are against it. At a small nune in an out-of-the-way district, as a temporary method it is worth trying.
When concentrate to be treated by method (2) has been weathered, it becomes highly acid, and needs special alkaline washes.
According to method (3) all complex concentrates, that is, those containii^ copper, zinc, or lead, or complicating impurities, should be smelted, although in Western-Australia pyrite containing as much as 21 per cent arsenic has been roasted and cyanided with a 90 per cent extraction.
Cblorination or method (4) is not much used now. At Bendigo, Victoria, Australia, one custom " pyrite " works mes chlorine in the vats, and this pyrite contains as much as 12 per cent arsenic.
Referring to method (5) the concentrate plant of the Treadwell group of mines uses fine grinding and agitation with cyanide solution. At Waihi, Western Aiistralia, a concentrate treatment plant specially treated 5368 tons in 1911 equal to 1.42 per cent of the ore crushed at the mills. This concentrate product contained 5.25 oz. in gold and 3.2 oz. of silver per ton, and the recoveries were 96.2 per cent of the gold and 94.8 per cent of the silver, at a cost of 16 per ton of concentrates treated.
As to method (6), at many places the concentrate is roasted prior to cyanidation, especially at the Goldfieid Consolidated, Goldfield, Nevada, and at Kalgoorlie, Western Australia. At the Ivanhoe mine at the latter place, the annual output to be treated is 24,000 tons, which is collected in ordinary tanks roast«d in five Edwards furnaces, Fig. 74, mixed with cyanide solution, ground in five 5-ft. pans. Fig. 135, agitated and filter-pressed. An extraction of 95 per cent was obtained at a cost <^ $2.68 per ton.
The five mills on Douglas Island, Alaska, contain a total of 900 stamps, and crush approximately 5000 tons per day. The crushed ore after amalgamation is concentrated on 360 Frue vannera, yielding an average of 90 tons of concentrate per day, of from 2.5 to 4 oz. gold per ton. To treat this product, a cyanide mill, owned jointly by the Alasksr-Treadwell Gold Mining Company, the Alaska Mexican Gold Mining Company, and the Alaska United Gold Mining Company, was built to treat the concentrate made in their various mills, and has proved an imqualified success.
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The Alaska-Treadwell Mill
THB ALASKA-TREADWELL CONCENTRATE TRBATHENT-n.AHT
coneistB of three buildings situated on a hillsde 200 ft. above the stamp-mill. The upper building contains the grinding taid amalgamating plant, frith a lower floor for solution-storage tanks. The lower contains the cyanide equipment proper, -while the refinery is in a concrete building at one aide. The flow-eheet of operations is shown in Fig. 133.
The concentrate is received in two lOO-ton steel storage bins, 4, 4, 15 ft. diameter, with 55° conical bottoms Here it is kept covered with water, which effectually prevents oxidation of the sulphides. From this point until the cyanide treatmrait b^pns, the concentrate is in strong lime solution at all times. At the apex of the conical bottom of each bin, tight-fitting gates control the outflow, which is at once sluiced directly into Dorr
Fia. 132.— Outline of Tube Mill Circuit, Aluka-Tnttdwell Mill.
clasaifieis, 6, 5. The sluicing medium is the coarae return product referred to later. There are two Dorr classifiers driven by one 7.5 H.F. electric motor, one feeding into each tube-mill and making twenty-four strokes per minute. This rate of speed (causing greater agitation) was found necessary to separate the large bulk of the fine from the coarse.
The coarse product (rf the classifiers falls into the spiral feeders of the tube^nills. These mills, 6, 6, are of the Abb^ type, 5 by 22 ft., with corrugated sectional liners: and 3-in. Danish flint pebbles are used for the grinding. Am^gamation was formerly part oi the process, but the whole product is now being cyanided direct without this. From a sump in the launder, an air-lift elevates the pulp to a spitzlutte, from which the coarse material is continuously drawn into a Dorr classifier, 11, the coarse from which feeds a 4 1^ 12-ft. Abb6 tube-mill, 12, similar to the larger ones
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Treatment Of Gold Mill Concentrates
described above. The dischai^ from this mill joins the overflow {ram the spitzhitte, and is elevated by air-Ufbi to two settling-cones, so situated that the ^igotniiBchai^ from them becomes the sluicing medium for the ori^nal feed referred to above. The overflow from the Dorr classifies passes into two Callow dewatering cones, the spigot product of which flows into latmdere, thence into a 6-in, pipe, 37 ft. long, having a fall of j in. per foot, which conveys the pulp directly to the lower or cyanide building. In the lower building the pulp is received into a wooden dietnbuting-box, from which it flows into four Wt. Callow cones. The spigot-product from these cones discharges into four similar ones placed lower than the first set.
The sp^t-product fnwn the lower cones enters one of four Pachuca tanks, 22, where it receives a preliminary treatment of three hours' agitation in a solution containing 2 lb. of lime per ton (0.1 per cent), after which it is allowed to settle and the clear solution is decanted. The filling, agitating, settling, decanting, and discharging of a 26-ton charge ot concentrate, which includes 46 tons of lime scdution, requires somewhat less than twenty-four hours. This preliminary treatment saves in the subsequent treatment at least 1 lb. of cyanide per ton of concentrate. The overflow lime water from the Callow cones enters the same sump with the decanted lime-water from the prelimi- — mm~...!B. nary treatment, and is pumped ^Sliir'^ into a reservoir of 75 tons' ■^tiSS^Si^' capacity atuated in the upper building. The thickened pulp, ranging from 1.8 to 2.2 specific gmvity, is drawn into one of eight Pachuca agitation tanks, 24, where it is given the cyanide treatment. AU Pachuca tanks in the mill are 10 ft. in diameter and 30 ft. high, with 60° conical bottoms. When filled to iJie level found best for agitating (which is 6 in. below the top of the central odumn), each tank h(ddB a volume equivalent to SO tma of water.
Fio. 133— Flow-tiheet. .'Maeka Treadwell Mill
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The Alaska-Treadwell Mill 217
This is equal to the regular chai^ of 30 tons of concentrate with 40 tons of solution. The floors under the Pachuca tanks, as well as all other floors in the building, are <^ smooth concrete, sloping to a central sump, supplied with smalt piunps to return any escaped solution and to pump it to the proper tanks.
The first cyanide treatment consists of eight houis' agitation in a 2-Ib. (0.1 per cent) cyanide solution, either potassium or the mixed cyanides being successfully used. Alkali is kept at 1.25 lb. (0.063 per cent) of lime (CaO) per ton of solution. Lime is added during the treatment if the titrations show below that figure ; eighteen hours is allowed for settlement and decantation of this solution. Decantation takes place through a flexible hose.
The long settlement allowed, with the excessively fine condition of the concentrate, its h^ specific gravity, from 4.6 to 5.0, and the high alkalinity of the solution, leaves a 30-ton packed mass in the bottom erf the Pachuca. This is brought into Station within fifteen minutes by a device desi^iated as the " spider," which is an adjustable hollow a-Tinnlftr casting with radiating fingers, the whole encircling the central agitation-column, see Kg. 100.
The second cyanide tieataient of the charges is with solution drawn from the barren-eolutitm storage tanks or the wash-solution storage, the cyanide strength being 1.5 lb. (0.076 per cent) per ton of solution. After two hours' agitation the air is shut off and almost immediately decantation is started. This decanted solution is pumped directly on to an incoming fresh chai^, being strengthened in cyanide as it enters the tank, and becoming the first cyanide solution for the new charge. This cycle in handling solution — barren to wash-solution, then to second cyanide treatment at 0.075 per cent cyanide, then to first treatment at 0.1 per cent cyanide, thence to precipitation and back to barren — gives at each step, just the conditions best suited for that step, and is very satisfactory in practical operation. The settled pulp after the second decantation has a specific gravity of 1.8, and is readily agitated by means of the spider, and then discharged into the pulp-storage tank by a Byron Jackson 4-in. centrifugal pump.
Filtering is done in two type-l-B Kelly presses. By opening valves in the circulation-lines directly under each press it is fiUed with either pulp or wash-eolution as desired. The excess pulp or wash-solution from the press-cylinder is returned into its proper line by displacing with compressed air admitted into the cyUnder. The amount of wash given depends upon the comminution of the concentrate, the usual pulp being washed with 0.5 ton of solution per ton of concentrate. The cake formed during decantation of the first treatment-solution, being very fine slime and more impervious to wash-solution than the regular pulp, is given 1 ton of wash
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218 Treatment Op Gold Mill Concentrates
per ton of concentrate. When filling the pieBs, the contained air is allowed to escape through an overhead pipe attached to the highest point <rf the press-cylinder. The change in sound of the exhaust indicates to the pressman when the press is full. After drying the cake with compressed air until it contains not more than 10 per cent of moisture, the press is opened and the cakes shaken <^ with wooden paddles, and then shiiced with water to the tailing-dam. A distributer below the press-launder sends the gold-eohition to two gold-sumps and the waab-eolutioii to the two wash-sohitioD storage-tanks. These four tanks, as well as a darifyingtank which is in the same group, are built of 3-iu. redwood, 15 ft. in diameter by 16 ft. deep, aud each holds 75 tons of solution.
The wash-solution is pumped to a Fachuca tank as needed, becoming a second-treatment solution. From the gold-tank the solution is drawn into the clarifying-tank, in which are suspended vertically six canvas filter-leaves, all coDuected to the suction of a triplex 7 by 9^. Aldrich electric pump, used exclusively for pumping gold-solution through die precipitation-presses. A traveling-belt, driven by ratchet^ears and a pair of eccentrics connected to the pump-drive, feeds zinc-dust into a c<me. Here the dust is emulsified with a small stream of gold-solution tapped from the dischai^^e-column of the same pump, and is then drawn into the suction-line. An automatic float in the cone prevents the introduction of air into the pump-suction. The pump raises the solution with the zinc dust to the upper part of the building and forces it through two 36-in. triangular, l&-frame Merrill presses. An average of 145 tons of solution is precipitated daily, with a consumption of 1-3 lb. of zinc dust per ton of solutitm, equivalent to 0.86 lb. of zinc dust per ton of concentrate. The average strength of solution before precipitation is 1.25 lb. (0.0626 per cent) of cyanide; 1 lb. (0.05 per cent) of lime, and $9.50 (0.2 dwt.) gold. The barren or precipitated solutions are kept at 10 cents (2.3 grains), or leas, gold per ton, and are used for wash-solution or returned to the Pachuea tanks, as desired.
CONCEIfTRATB-TRBATHEHT AT THB OOLDFKLD CONSOLIDATED HILL OOLDFIELD, IfEV.
The raw concentrate amounting to 6 per cent of the weight and con-tuning 67 per cent of the value of the ore, is collected in flat-bottomed-agitator tanks. It is here neutralized with lime and pumped to three Fachuca a^tators, in which It is agitated during eight-hour periods in a 2-lb. solution of cyanide. Decantation at the end of the period is still practiced and the charge is re-agitated with a freshly precipitated solution. Five periods of eight hours each, followed by decantation, are Etufficient to remove from SO to 85 per cent of the value of the concentrate. It is the intention to s^id to the roaster a product valued at $25 to $30, and this
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Goldfield Conc. Mill 219
treatment is varied with the grade of the ore so as to accomplish this remit. The pulp from the Pachucas, when dissolution is completed, is delivered to a storage tank from which it ia pumped to Kelly filter-press for filtration and dryii^. Thia drying ia accompliahed with air and the moisture ia reduced to 12 per cent. The consumption of cyanide during the raw treatment ia 2.5 lb. per ton and lead acetate ia uaed in the {vopoitum of 1 lb. per ton of concentrate.
The product is diunped into a bin, and a 14-in. conveyor, set at an an^e of 17°, carries it over a Blake-Denniaon automatic weighing machine en route to the bina m the roasting plant. The concentrate from this conveyor ifl distributed by means of a swinging bucket elevator to two bins , having 45° sloping bottoms and 1620-cu. ft. capacity, from which it is fed by means of two 12-in. screw-conveyors, making three-quarters of a revolution per minute, to two slow-^noving belts. These belts discbarge the concentrate through the arches of the furnace between the first two rabbles. Roasting of concentrates is done in two 54-Bpindle duplex Edwards furnaces, each having 1456 sq. ft. hearth area. The capacity of each furnace ia 40 tona of ccmcentrates per day, although the amount roasted is the two furnaces is approximately 55 tona per day. The raw concentrates, after a preliminary cyanide treatment, assay 1.23 oz. Au and 18.76 per cent S, the sulphur, after roasting, being reduced to 0.90 per cent. The coat of roasting, per ton of concentrate, is 10.82; while the ccanplete cost of the two furnaces, dust flues, stack, 65 by 158 ft. steel building and miscellaneous bins and machinery amounted to $70,459.16. The concentrate loses 17 per cent of its weight in roasting; and of the 1) per cent (^ the material passing out of the furnace as dust, only ) per cent is lost out of the stack. Five Merton furnaces for the roasting of 120 tons per day of Kalgoorlie aulpho-telluride ore cost $38,900.
The bins are filled on the day shift, and have sufficient capacity to run for twenty-^our houis. Each furnace requires 4} H.P. By means of iron goose-neck fines, the gases from the roasters at a temperature of 450° F. are delivered to a concrete dust-fiue 264 ft. long, having a crosa-section of 50 sq. ft. From thia flue, 20,700 cu. ft. of gases per minute escape through a steel stack, 100 ft. high and 54 in. diameter having a temperature at the base of the stack of 325° F. Velocity of the ^sea in the dust-flue is 7^ ft. per second.
The roasted ore is discharged into a Baker cooler, 5 by 22 ft., revolving in water with about 40 per cent submergence. It is delivered to one tank for twenty-^our hoxira, then settled and decanted to a consistence of 1 to 1 and sulphuric acid added in the proportion of 20 lb. per ton of concentrate. Agitation with the sulphuric acid is continued for eight hours. Water is then added to fill the tank and the charge allowed to settle. When clear, the wash is decanted and the tank refilled with fresh
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220 TREATMENT OF GOLD Mill, CONCENTRATES
water. Four water washes are given, equivalent to eight tons of wash water per Usi of ctmoentrate. All washes are clarified and the overflow sent to ax redwood tanks, 10 ft. diameter and 5 ft. high, arranged in series for recovering the copper. These tanks are kept filled wiUi cyanide tins and all kinds of scrap from the mill. The average copper content of the washes is 0.4 lb. per ton, and 70 per cent is recovered.
The thoroughly washed charge is neutralised with lime, and by means of centrifugal pumps elevated to one of four Pachuca agitators, 14 ft. diameter by 25^ ft. high. Here the roasted charge is agitated for ei^t hours in a 2-lb. solution of cyanide, containing 1.2 lb. CaO as protective alkali. At the end of e^t hours, agitation is discontinued, the charge settled, decanted, and re-agitated with a freshly precipitated solution in the same manner as described above in the treatment of the raw concentrate. Five periods of agitation followed by decantation are given, and a total of 3 tons of solution per ton of concentrate is decanted. Consumption of chemicala amounts to 4) lb. cyanide and 2 lb. lead acetate per ton of concentrate. After agitation is completed, the settled charge is delivered to a storage tank 18 ft. diameter by 8 ft. high, fitted with the adjustable square-shaft agitator. Placed centrally in the bottom of this tank is a 4-f t. cone with pipe connections through which the thickened pulp is fed to a 5 by 18-ft. tube-mill. The pulp issuing from the tube-mill is elevated by means of a belt and bucket elevator back to the above-mentioned storage tank. This circulation grinding is continued for sixteen hours, at the end (rf which time 95 per cent of the material will pass a 200-mesh screen. From 80 cents to $1.25 per ton is removed in this circuit. Since the change uf solution increases extraction and since the final tailing is sent to the mill proper for filtration, it was decided to re-grind after the greater part of the gold had been removed. After re-grinding, the pump is delivered by means of a centrifugal pulp to the filter stor^e tank in the milt proper, mixed with the mill pxilp, filtered, and sent to waste. Costs are as follows; Labor, S1.02; power, 10.78; and total supplies, S3.88; making a total cost of $5.68 per ton. During 1912 results were as follows: value of raw concentrate, 6.58 oz. gold; value after treatment, 1.23 oz. gold; recovery, 81-3 per cent; value of tailing after roasting and treating 0.097 oz. gold; recovery, 92.16 per cent; total recovered from roasted material, 17.23 per cent; and recovered from both treatments, 98.53 per cent.
The treatment at each plant was as follows:
(A) Raw concentrates are agitated in 2 lb. KCN solution for five e^t- hour periods, decanting after each period; after filter-pressing, they are roasted in two Edwajrds duplex 54-«pind1e furnaces. The roasted concentrates are agitated for eight hours in H2S0t solution (20 lb. acid per ton of concentrates). After four water washes, the charge is neutraUzed ard agitated in 2 lbs. KCN for five eight-hour periods, decanting after each
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CYANIDING CONCENTRATES CTAHnDATION OF GOLD BEARING CONCEICTRATBS
r
5 i
Iba.
Coar ?■■ Ton or Cokc'ts.
nut.
Ao.
Ac
Au,
As.
1
1
1
t
1
1,
1
J— Goldfield,
Com, Nev...
B— AlMk»,Tread-
B.58 2.92 16.0 2,2
7.
K-6E
C-Geldenhuii Deep, Tnmsvul D— Oriental, Com, KoK».-.
Then it is tube-milled for sixteen hours and sent to the regular plant 61ter, using 3 lb. lead acetate per ton ol concentrates. (J. W. Hutchinson, Min. and Scientific Press, January 25; February 1, 1913.)
(B) Clean and docile pyritic concentrate is ground throu^ 200 mesh ; agitated in 1.5 lb. KCN for eight hours; decanted; agitated for four hours with barren solution; filter-pressed and precipitated by idnc dust. (1911 Mine Report, Min. and Scientific Press, June 29, 1912.)
(C) Mill scrap and black sands. After amalgamation, the concentrate pulp for cyanidation (98 per cent 200 meeh) assays 16 oz. Au. (R. Lindsay, J. C. M. A M, So., S. A.)
(D) All-filiming in a tube-mill with KCN, then forty-eight hours' combined air agitation and leaching, followed by filter-pressing.
ASration. — Ox^en-sbsorbing compounds, such as pyrrhotite, hornblende, etc., not only cause an increased cyanide consumption, but frequently reduce the extraction of the gold by robbing the solution of its oxygen. Suitable oxidation of the pulp, prior to the &ppUcation of cyanide solution, by forcing compressed air through the charge, render many of these oxide-consuming compounds harmless.
Roftsting. — In the treatment of ore in which the gold is intimately associated, either physically or chemically, with such telluride ccnnpounds aa sylvanite or calaverite, or with arsenical or antimcmy compounds, preliminary roastii^ is frequently the only known recourse. The practical object of the roast is to so liberate the gold as to permit ample contact with the cyanide solution and destroy deoxidizers and cyanicides. (A " dead " or " sweet " roast is usually essential.) The most extensive application of roasting is at Kalgoorlie, though some Cripple Creek ores receive
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222 Treatment Of Gold Mill Concentrates
this preliminary treatmrait, as do the graphitic ores of the Ashanti Goldfields. In the treatment of rebellious concentrates by cyanide, roasting plays an important part at the Goldfield, Nevada, plant.
The usual furnacee are the Edwards, Merton, Fearce, and HoIthofF.
On the teUuride ores of Kalgooriie, crushed through approximately 28-mesh roasting with Edwards and Merton fumacee, coste approximate 65 cents per ton of ore. The sulphur content of the raw ores varies from 3 to 6 per cent, although the elimination of the sulphur affords only a TOi:^ indication of the success of the roost. The hearth area, per ton of ore roasted per day, varies from 17 to 29 sq. ft. Consumption of wood varies from 10 to 13 per cent of the weight of the we. Tlie temperatiue is approximately 8fiO° C.
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Chapter Xviii
Various Treatments And Calculations
FLOTATION AHD CTAmDIHO AND CALCULATIONS
Concentntion by Flotatkm, Cysniding or Smelting ttie Concentrates and Cyaniding the Tailiags. — The method would be suited to a low-grade gold ore, that, owing to itfi refractory nature, could not be profitably treated by cyaniding.
First Ctse. — ^The ore would be aU-slimed, then subjected to flotation,; yielding a amaU proportion of high-^p^e flotation concentrate and a clean' tailii^ for cyaniding. The concentrate would be subjected to special cyanide treatment as described under head of " treatment of gold mill concentrate."
Second Case. — The ore would contain copper or zinc sulphides that would interfere with cyaniding. The ore would be crushed, saving the concentrate and wasting the tailings. The concentrates would then be shipped to the smelter.
Third Case. — ^Tbe ore could be cyaoided and the tailings concentrated. In this case the aim is to treat by cyaniding alone, but from the tfuUngs to recover some concentrates by means of flotation. The concentrate would be subjected to special cyanide treatment.
DKTinO AND CYANIDING
The ore of moderate grade contuns graphite that would interfere with cyaniding. The crushed ore would be dried at 150° C. so aa to render the graphite inactive and less flocculent. It would then be cyanided.
Tbeathsnt Of Tailings From Acid Or Ammonia Leachino
The original ore has a quartz gangue. If treated by sulphuric acid leaching it must have but little lime or magneoa carbonates. If treated by ammonia those carbonates do not interfere. In either case the ore carries oxidieed copper minerals and even some microscopic metallic copper. The copper having been removed the tailing is in fair state to reccnrer any contained gold or silver by cyaniding (see " Sulphuric Acid or Ammonia Leaching of Copper Ores.")
Calchlatioh Of Tohnaobs In Hills
In wetrcrushii^ mills, concentration and hydro^netallurgical works it is often desiraUe to measure the water and ore handled, either the amounts
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224 Various Treatments And Calculations
contained in tanks or the quantitiea fmesing in a given time. In some mills such measurements are systematically made, but in others the amounts are merely guessed, or they are measured once and ever after aasunted to lemun cmistant. Discrepancies between theoretical and actual recovery are due to errors in sampling and assaying the material before treatment, added to the corresponding errors affecting the material after treatment, and multiplied by errors in the estimate of the tonnage treated. The last item is therefore fully as important as the othera in calculating probable retumB.
The tonnage of sand in vats filled by settling under water is best ascertained by means of boxes of stout sheet iron (conveniently made of exactly 1 cu. ft. capacity, but in any case accurately measured) , having a nmnber of small perforations in the bottom and provided with handles. Several ot these are placed in the vat at various stages of the filling, and are allowed to remain throughout the treatment. While the vat is being discharged these are carefully removed and " struck " level; the contents are then dried and weighed, giving the pounds of dry solid per cuMc foot. Several charges should be thus tested and averaged to obtain a constant value for the ore or tailing treated. It is demrable to place some of these boxes near the center and others near the periphery of the vat, so as to represent variations in horizontal as well as vertical distribution. The mean weight per cubic foot and the volume of sand ia the charge ^ve the total wei^t ot sand. While cubic boxes are often used, a cylindrical form is preferable, as being less liable to deformation.
In estimatii^ the cubic content of a round vat several diameters should be measured (preferably three or four making eq\ial angles with each other at about the middle depth), and for the greatest accuracy a similar set of measurements should be made near the bottom and another near the top, the arithmetic meui of all the diameters measured being used in the computation. In supposedly cylindrical wooden vats of over 25 ft. diameter difTerences of 6 in. or more may be found, due to imperfect construction, to settling, or to unequal shrinkage of the staves, owing to their upper portions being intermittently dried while the lower ends remain wet.
If JD be the internal diameter, and H the depth, in feet, of a cylindrical tank, the volume is 0.7854 D^ cu. ft., 0.024544 D^H fluid tons, 5.89 WH U. S. gaUons, or 4.008 D^H imperial gallons.'
> The U. S. gallOD will hold 81 lb. of water, the Bridiih imperial KaUon 10 lb. ot walw.
The percentafe P, of dry dime in the pulp, is computed by the formula
where ^ is the sp. gr. of the dry elinio, uid a the ep. ff. of the wet pulp. The ^i. gr. ol the pulp cui be ascertained by a hydrometer or by weighing a unit meaaore id it. ,
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The capacity of a filter-preae is most accurately determined by blowing the chaj^ aa nearly dry ob possible before opening, then selecting a certain proportion of the frames at eq\ial distances from end to end of the press, weighing the entire content of each separately, and taking an individual moistuie sample from each frame tested. The dry wei^t cf slime in each is separately calculated and the average multiplied by the number of frames.
In ascertaining the wdgbt of solid in a vat filled with uniformly liquid pulp, such as slime in an agitator, a cubic foot or any convenient measured vohizae may be dried and the residue weighed, whence the weight in the entire volume is obtained by proportion. If the mixture is weighed before drying the percentage <A solid in tbe pulp may also be found.
A much easier and more rapid method is to find the specific gravity ot the pulp, either with a hydrometer or by weighing a liter or other convenient volume. The density of the dry atdid must also be known, at least approxiznatety. Knowing these two vahiea, specific gravity of mixture and density t4 dry solid, the weight of dry solid per cubic foot can at once be calculated.
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Chapter Xix
Smelting Of Gold Orbs
BLAST-FDRNACE SMBLTHTG VS. CTAHIDINO Ot GOLD ORBS
- Gold may be recovered from ita ore by the prooeaBes of alver-lead, or of copper-matte smelting. It <rften is found in copper-or lead-bearing oies and when in exc^s of 0.02 oz. per ton, is paid for t^ the smelting works at the rate of (19 to S19.50 per ounce. Practically all the gold is recovered in smelting, and this would be the best method of treatment were it not for the high cost of freight and for treatment. If smelted near the mine in a works operated by the mining company, the cost of frei^t is eliminated. The charge for lead-free, fairly siUcious ores, from Cripple Creek and from Boulder county, Colorado, is from $4 to {10 per ton, according to grade. The low-grade ores are subject to a low-treatment rate. On the other hand, ore treated by milling and amalgamation, or by cyanidation, while the extraction is leas, often yields higher net retunis. A sample is found in the case of the sUicious gold ore from Boulder coimty, Colorado, containing 0.5 oz. Au per ton, giving 70 per cent extraction by milling and amalgamation, or of 90 per cent by cyanidation. In comparing the costs we have:
8Xbi.Tim0
100 per cent erf 0.S OS. Au ftt S19.00 tO.SO
Mining 2 . 00
Freight 1 . 60
Treatment 4.00 7.50
Net returns. S2.00
70 per cent of 0.6 tw. Au at $20.50 $7.17
Mining 2.00
MUling 1 .00 3-00
Ctahidatiom
00 per cent of fl.5 oi. Au at $20.60 $9.23
Mining 2.00
Cyaniding 1.66 3.06
Net returns $5.«8
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Gold Ore Prices M7
From the above comparison it is seen that cyaniding is the meet profitable method of treatment for this grade of ore, and at this place.
Price at Gold Ores. — When lead free or so-called diy ores cinitaining gold (and silver) are sold to a smelting works they are paid for on the basis of dry ore, which see under head of " Purchase of ores, silver-lead smelting."
Costs at du Bdmont Tonoiwh Gold HiU in 1914-191S.~This mUl of SOO tens daily capacity has, per ton of ore put through, labor, $0,419; supplies, $1,318; power, $0,419, using 1.68 H.P. per ton; being a total of $2,156 per ton of daily capacity.
Costs at ttie Homestate Gold Mill in 191S.— Cost at stamp-milling and amalgamating, $0.2811 per ton of ore. Hie tailing from the mill were regroimd at a cost in 1914 of $0.1264 per ton of product reground. By classification, this reground material yielded two products, sand and slime. The sand was leached in vats 44 ft. diameter at a cost of $0.1772 per ton. The other product, the slime, was all filter-pressed and all the slime plant operating costs were $0.1838 per ton in 1914.
Costs at the Moddeifontein Gold MUl.— Rand district. South Africa, in 1914, the cost was $0,686 per ton of ore milled.
Speaking broadly the treatment cost per ton of concentrate will vary between $2.50 and $5 per ton, depending upon tonnage cost of supplies delivered at the plant, also labor. The extraction of gold, using the allsliming process varies between 90 and 97 per cent on a raw concentrate amenable to cyaniding, as at the Oriental Cons., the Atasko-Treadwell, the Esperanza, Waihi and elsewhere.
The price ot gold as sold to the mints is unchan^g, being $20.67 per troy ounce, 1000 fine. From this the mint makes a deduction of 2 cents per ounce to cover the cost of melting and assaying.
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Part Iii Silver
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Chapter Xx
^LVBR, ITS ORSS AND THEIR TRBATUmTT
Phyaical Propertiea of SOver. — ^This is the whitest of metais, huder than gold, softer than copper, more malleable and ductile than any but gold, and the best of conductors of heat and electricity. Its specific gravity is 10.5; it melts at 962° C. and boils at 1850° C, then volatilizing and yielding a green vapor. When pure and molten it will absorb oxygen, which when the metal again soUdifies causes the so-called putting fd the metal, well known to assayers.
Charactbkistics Of Silver 0Kb8
The silver minerals of importance in treatment are aa follows:
Native silver, which sometimes occurs as flakes or leaves, and as wiie- Eolver and metallic silver adherent to native copper. Native silver can be readily amalgamated, but when present in particles of visit:^ size it is so slowly soluble in cyanide, that practically no extraction can be obtained.
Cerai^yrite (horn-silver, silver chloride), AgCl, is widely distributed. At mines it is found in the upper oncHzed zones; It is probable that much of the so-called chloride ore is r6ally a cbloro-broipide (embt^te). The ore is readily 'amalgamated and is free-^nilling. The silver chloride of it also ia readily soluble in cyanideand in sodium hyposulphite solutions.
Argentite, AgxS, is one of the common silver ores. By using chemicals (bluestone and salt) it can be amalgamated in pans, and the silver extracted thus from the ore. It ie soluble in potassium cyanide solution.
Stephanite, 6Ag2S,Sb3Ss; pyrargyrite, SAgiSgSbaSs; proustite, SAgaSrAssSs; drycroaeite, AgaSb, are silver sulph-aisenides or sulphantimonides, refractory in amalgamation, even with chemicals, sparingly s(Juble in cyanide solutitm, but readily soluble in a solution of merourous potasmc cyanide.
Finally we have those silver sulphides that contain also copper. These are polybaate, 9(Ag2Cu)S(SbAs)2Ss and tetrahedrite (gray copper ore, tahlerz), 4CuFeAg2(HgZn)S,(SbA9)Sa, the most complex of all, in which the ffllver varies from 0.06 to 31 percent, being higher in the. arsenical and bwer in the antimonial varieties. These sulphides are refractory to any amalgamation method and because of their copper content are precluded
232 Silver, Its Ores And Their Treatment
from treatment by cyanide, even when roasted. This does not mterfere with treatment by hyposulphite Uxiviation after roasting.
A number of rare minerals containing silver could oIbo be enumerated, but for tbe metaltur^st the minertds above named are the important ones.
Klver ores in general contain but a small percentage of precious metal, lliey are composed mostly of gangue (waste matter of the ore) and many are treated that contain lees than 0.1 to 0.2 per cent silver. Thus we have at the Comstock Lode, Nevada, silver in native form and as sulphide, but oxides of iron and manganese with the associated sulphides, pyrite, blende, galena, and chalcopyiite. At the Ontario mine. Park City. Utah, the silver occurs as argentite and tetr&hediite in a gangue of quartz and clay associated with a little of the heavy minerals blende and galena. These sulphides carry diver which is recovered with the concentrate in case of concentration.
THE EXTRACnOII OF SILVER FROM ORES
Silver is extracted from its ores by milling methods, and by smelting. Certain ores contain the silver in a form suitable for cyaniding, and in consequence that method of treatment is coming forward. The other methods have largely dropped out of use. No reason appears why hyposulphite Uxiviation should not revive under the stimulus of the recent methods of agitation and filter-pressing. The patio process formerly much practiced in Mexico where ixHiditions favored, has been superseded by cyaniding in many cases, on account of tbe lower cost of operating tbe latter process, but in the past, la^e quantities of silver have been extracted by tbe patio process.
Txrathbht Op Silver Orbs
These, as in the case of gold ores, may be treated by milling (m- 1^ smelting. Milling processes, with the except<»i of amalgamation, ate called hydrometalliu^cal methods. Since these methods are often c<Ha- Inned with amalgamation and concentration it would appear that ' all might better be grouped under head of silver milling. We may divide the methods of silver milting into:
A. Amalgamation.
(1) Wet enhxT-miMng, or the Washoe process.
(2) Plate and pan amalgamation and concentration.
(3) Dry aihier-miUing, or the Reese River process.
B. Mining using hydrometallwgical processes.
(1) The Avfputin process, based upon the 6<^ability of silver chltmde in tnine.
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Silver Ore Treatments 233
(2) The Ziermyel proeeaa, dependent on the sohilnlity of silver sulphate in hot water.
(3) The Patera -procett, in which silver chloride disBolvefl in a sohitioD <tf sodium hyposulphite.
(4) The RunuH proceaa — a modification of the Patera process in which a so-called " extra solution " is used.
(5) The cyanide proeeaa, in which the olver minerals either with fX without roasting, dissolve in dilute cyanide solution.
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Chapter Xxi Amalgamation Of Silver Ores
The silver ores suitable to treat by milling and amalgamation are thoee that contun the metal in such form ua to be acted upon by mercury when assisted by agitation, heat, and certain chemicals. The ore is first crushed fine by stamps, af< in gold milling, then treated for several hours in grinding pans, the reactions being slow compared with those of the amalgamation of gold. In gold pulling, the greater part of the gold can be arrested on ao apron-piste during the few seconds in which the ore is passing over it, while in silver milling, the ore-pulp has several hours' contact with mercury, uded by heat and chemicals, and is but slowly amalgamated. In gold milling, ore containing 0.5 oz. Au per ton can be profitably milled. In silver milling, ore of equivalent value would contain 10 ot. Ag per ton, or 20 times as much metal. Thus is seen why ao much time is allowed in mlver milling, and why bo many precautions must be taken to be sure that all metal possible is recovered. Several ounces of solver per ton often remain in the tailing.
The silver metals suited to pan amalgamation are cerargyrite (hom-silver, silver chloride), native silver in flakes, wire, or other forms, and certain silver sulphides, notably argentite (AgaS). When the ore is refractory, containing areenicat and antimonial sulphides, and especially containing tetrahedrite, galena, or blende, it is necessary to roast with salt, setting free the mlver or converting it into the form of a chloride, which becomes susceptible to anudgamation. There is no sharp line of demarkation between free-milling and roasting-milUng ores. Often the upper part of a vein is free-milling while in depth base metals and sulphides begin to come in, trad it finally becomes necessary to roast the ore. The best extraction therefore is obtained from decomposed or oxidized ore, in which the ulver materials occur in a form that renders possible the action of the mercury. There are few deposits of oxidized ores containing alver chloride and native silver that as a whole are suitable for free silver milling. Such ore, so far as silver chloride is concerned, can also be treated by cyanidation, but the latter method would not recover native silver.
Arsenic and antimony compounds interfere with amalgamation by fouling the quicksilver, checking the reactions of the chemicals added to 234 .
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Wet-Silver Milling 236
promote amalgamation, and by carrying off silver, which is incapable of being amalgamated with them.
The Waslioe process, developed throi^b the combination of the California stamp mill with an elaboration of the Norwegian Tina for fine grinding and amalgamation (pan) and the chemicals of the Patio process, was introduced for the treatinent of Comstock ores in 1860. This amalgamation process, adapted to American conditions, rapidly assumed the same position in the United States for the treatment of Eolver-gold ores OS was occupied by the patio proceas in Mexico for the treatment of Hia same class of ores.
A later development when treating complex ores was to give a chloridizing roast preceding grinding and amalgamation in the pans. This was known as the Reese River ^ttcess.
(1) WET Sn-VSR-HILLIITO WITH TAHK-SBTTLINO
This is also known as the Washoe proce^ receiving the name from the place where it was perfected for the treatp.ent of ores from the Comstock Lode, Nevada. The process is applicable to the so-called free-milling ores, in which the sUver occurs native, as chloriae or in small amount as argentite. The ore should be free from lead and from any tough clayey gangue.
In wet silver-milling, the process consists in coarse-crushing the ore, stamping it fine, and collecting it in settling-tanks. The crushed sand is ground in amalgamating-pans using mercury to collect tiie silver: The sand is separated from the silver-bearing mercury in sett^g-pans, and is rejected. The amalgam is strametl from tha mercury, retorted, ittbd the retcfft-residue melted into silver ingots. Gold present in tiie ore is recovered as well as the silver. The process resembles gold-milling except that amalgamation and the removal of the amalgam is effected in pans.
fig. 134 is a sectional elevation of a wet-crushing tank-mill for the treatment of free-milling silver ores. The ore from the mine is amalgamated (H) plates precisely as in gold millie^, which see. The coaiBe crushing is done during the ten-^6ur day-shift.
Water (6 to 8 tons per ton of ore) is at the same time supplied in the mortar, and forms a pulp, which is splashed through the SO-mesh screens by tile motion of the stamps using a double-discharge mortar. The large screen-opening possible with a double discharge favors a more rapid pulverisation than would be possible with a single-diflcharge mortar. The pulp flows l^ launders / into the settling-boxes or tanks g 7 ft. square by 3 ft. deep. There is a double row of these tanks, twenty in a row, occupying the length of the mill in front of the stamps. The Sow of t^ pulp is from box to box in series, until it goes by launder to a settling-pond outside the mill. Moet of tiie solids settle in the first boxes, a-further portion
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Amalgamation Of Silver Ores
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Wet Silver Milling 237
dropping in the succeeding ones, and the turbid water passing to the pond. Here it has its final chance to settle before running to waste, or it may be again iised in the mill if water is so scarce that it pays to do this. The settled slime is dug from the pond at a later time and treated like the rest of the crushed ore.
A variation of this method, shown in Fig. 134, consists in conducting the Qow from the last box j;' by an inchned elevator to a tank h situated in front of and above the battery, the dirty water being again used for stamping. When the first settling-box is full the flow of the pulp is bypassed into the next one. The contents of the full box arc shoveled upon the floor adjoining, and thence taken as needed to the amalgamating-pans q. The emptied box has the flow of the last one turned into it, thus making it the last in the series, and the launders are so array^d that this can be done.
The ore, thrown out upon the floor, is fed directly into the pans or loaded into the tram-car seen in Fig. 134 and conveyed to them.
Fig. 135 represents a pan. It is 5 ft. diameter by 30 in. deep, and is furnished with a central sleeve or cone through which rises a shaft carrying a cylindrical casting called a B(nder, which becomes bell-shaped and broadens into feet below. The spider carries, bolted to the feet, a flat cast-iron ring called a muller, and to the under side of the muUer are attached six shoes or plates of chilled cast-iron 2i in. thick. The spider, muller, and shoes are raised or lowered as de- Mred, by means of a hand-wheel and screw at the top of „ ,„ „. ,
the shaft, which is driven by ^"'' ^35.-Piv.^foot Cont.nuom Gnndmg Pan. bevel gearing from the horizontal shaft and pulley below. Upon the bottom of the pan rest chilled cast-iron plates or dies that fumish the lower or fixed grinding surface. The shoes attached to the muller revolve M R.P.M. and rubbing upon the dies, grind the ore.
In working the pans, the shoes are raised ^ in. from the dies and set in motion, the pan is partly filled with water, and 3000 lb. of the damp pulverized ore is shoveled in. The ore and water nearly fill the pan and the mixture is stirred until it is of the consistence of honey. The motion estab-
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238 Amalgamation Of Silver Ores
ashes ft movement or current of pulp beneath the miiller toward the poii^ ery. At the periphery it rises, flows toward the center, anlcs, and passes again under the shoes. To assist the action, the rising pulp is deflected inward by cast-iron wing-plates.
After thorough mudng in the pan the shoes are lowered until they touch the dies, and grinding goes on for 1) hours, the content of the pan bong meanwhile heated nearly to boiling by steam under pressure from a pipe that dips beneath the surface of the charge, the pan being covered.
After grinding, the shoes are raised and 300 lb. (rf mercury (10 per cent the weight of the ore) is added, by sprinkling it through a fine strainer. He mixing is then ctmtinued four hours. The mercury takes up silver most rapidly at first, but the action afterward slackens. The globules oi mercury suspended in the pulp take the silver as they come in contact with it. Care is taken to have the pulp of the right consistence so that mercury will not settle out. This condition is shown when a wooden stick, dipped in the pulp and withdrawn, is found to be covered with a thick mud in which are disseminated minute globules of mercury. If the ore is refractory, salt and copper sulphate are advantageously added at the beginning of grinding to accelerate the reactions, promote amalgamation, and increase the yield of silver.
The charge above treated having been amalgamated, the pan is ready to empty into the settler r. About fifteen minutes before the dischai^ng, the speed of the muller is reduced to 40 R.F.M. and the pan filled to the top with water. A plug closing the dischai^ opening at the bottom of the pan, seen at the left in section, Fig. 136, is pulled out, and tbe entire content run by laimder to an 8-ft. settler, at a lower level, shown at the right of the amalgamating-pans in Fig. 134. Emptying the pan and washing it with a hose takes half an hour, after which time the plug is replaced, and tbe pan is ready for another charge. Thus the total time for the cycle of operations described is six hours, making it possible to treat four charges daily.
The reactions that take place in the pan are as follows:
Native silver m threads, films, fiakes, or grains readily combines with the mercury and forms an amalgam which contains a large excess of mer^ cury.
Silver chloride in contact with the mercury decomposes as follows:
The metallic silver liberated amalgamates with additional mercury. The particles oi iron, abraided from the stamps and the bottom of the pan, decompose the mercury salt and liberate the mercury as foUows:
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Wet Silver Milung 239
Many so-called free-millmg silver ores contain argmtite which in part is decompoaed by mercury as follows:
The sulphide t^ mercury thus formed is lost. We have already stated that cbemicalfi, notably capper sulphate and common salt, are added to promote the decomposition of the silver sulphide. There is added in Uie amaU gamating-pan from 6 to 18 lb. salt and from 3 to 9 lb. copper sulphate per. ton of ore treated. Tlie reactions as generally given are the following:
The chloride oS copper acting tm the silver sulphide decomposes it:
The silver chloride amalgamates as shown by reaction (1).
The complete separation of the mercury with the silver-amalgam is effected in the settler, there being one settler provided for two amalgamating-pans. The settler is 8 ft. diameter by 3 ft. deep, three times the capacity of the amalgamating pan, but of similar construction, as shown in Fig. 136. No grinding is required, but the pulp must be agitated with the wooden shoes with which the settler is provided. The shoes nearly touch the bottom of the settler, the exact height being adjustable. The grooved border at the bottom just within the sides of the settler has a slight grade to the outlet and mercury-well at the left. The mercury settles from the pulp, Sows to the lowest point and stands at a height that balances the hydroetatic head of the content of the pen. Since the specific gravity of mercury is 14 and the content of the settler approximately 1.5, the height of the merciuy is a little less than 4 in. The bottom outlet-hole of the well in plugged. At different heights in the side of the pan there are provided openings that are kept dosed by phigs. When the plugs are withdrawn the tailing and water, free from mercury, pass out of the pan.
The shoes of the settler having been set in motion, at the rate of 15 R.P.M., and raised 8 in. above the bottom, the contents of the two pans are run in, as has been described. Water is then added to within 6 in. of the top, greatly thinning the pulp, and filling the settler. After half an hour the shoes are gradually lowered until, at the end of two houra, they nearly touch the bottom. The purpose of the agitation is to keep the lighter portion of the ore (now called the tailing) in suspension, while the ^Iver-bearing mercury, the heavier particles of sulphide, and the particles of iron fnmi the stamps collect at the bottom. The stirrii^ is ctmtinued H hours, after which the behest plug in the side of the settler is removed, and the turUd water containing tailing is allowed to escape by latmder, a stieam of clear water being meanwhile allowed to flow thtough. l^e
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240 Amalgamation Of Silver Ores
plugs are then withdrawn one by one until the settler is emptied of all the content except the heavy portion containing sulphide, iron particles, and the mercury. Emptying takes half an hoxir, and the cycle of operations becomes six hours as in the case of the amalgamating pan. Since escaping tailing contains sulphide, it may be run over rifQes, or blanket-lined launders, before running to waste.
The silver-bearing mercury or diluted amalgam, a mixture of silver^ amalgam and mercury, collecting in the mercury well, overflows by an
Fia. 136.— EighUfoot Settler.
escape-opening indicated in Fig. 136. From the opening it paeees by a half-inch pipe to the amalgam safe shown at the right of the settler, Fig. 134. The safe, arranged to prevent theft of the ama^^am, is shown on a larger scale iu Fig. 137. The amalgam and mercury enter a conical canvas sack or filter. The mercury oozes throw^h the pores of the canvas while the amalgam contuning as little as 14 per cent silver is retained. Occasionally, after amalgam has accumulated, the sack is squeezed between the hands to remove the surplus mercury, and the compressed amalgam containing 20
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Treatment Of Silver Amalgam 241
to 28 per cent silver, is reserved for retorting. The mercury flows out at the bottom through an outlet provided, as seen in Fig. 137, and is collected at a lower level in the boot w, Fig. 134, erf the mercury elevator, shown at the right. The elevator discharges to a mercury tank 8 commanding the amalgamating pans, to which it is delivered as needed through the pipe shown in the figure. Over the stamps and the pans are seen the overhead tracks that cany crawls by which the heavy parts of the machines are lifted or transferred. This facilitates the work of repairs and replacements.
The loss of mercury is commonly 1 to 1.51b. per ton of ore treated. A part is lost in handling, but the principal loss is the flouring, which causes the mercury to escape in the tailing. The loss is greater with talcose or clayey ores, and in those carrying cerussite, chalcopyrite, or galena. Loss is caused by grease coating the particles of mercury, in case this enters the ore from the machineiy. Ro. 137.-Aimlgam Safe.
Treatment of the Amalgam. — Since the weight of metal recovered in silver milling is much greater than in gold milling, the retorting of amalgam must be performed on a larger scale. Fig. 138 shows a sectional elevation and a plan of a combined retorting and melting furnace with the overhead crawl and chain-blocks by which the large melting crucibles are lifted from the fire in the melting furnace and transferred for pouring. At the left is shown in the elevation a cross-section of the cast-iron cylindrical retort which is 10 in. diameter inside by 28 in. long, resting upon arched cast-iron supports. There is a horizontal pipe, and a vertical water-cooled pipe, not shown in the illustration, in which the mercury condenses and from which it falls into a tub of water below. As seen in the plan, the front end of the retort is provided with a cover which can be securely clamped in position.
The charge of amalgam, containing 20 per cent mercury, should weigh 600 lb. and only half fill the retort. After filling, the cover is clamped on, first luting the joint with flour paste. A wood fire is started on the grate under the retort. The temperature is kept low at first, increasing to a red-heat at the end, i to } cord of wood being used. The operation lasts ten to fourteen hours, care being taken not to heat the retort rapidly, nor, for fear of blbtering it, to raise the temperature too high. The fire is then allowed to bum down, and the retort to cool. The lid is taken off
242 Amalgamation Of Silver Ores
and l^e silver residue removed to a crucible. This is seized by basket-tongs, which clasp it finely so that it can be lifted by the chain-hoist, transferred by the crawl to the ingot mold, and poured. These molds, 11 in. long by 4) in. wide and deep, hold 1000 oz., or 70 lb. silver.
Fici. 138. — HorizoDtal Retort and Melting Fum&oe Tor Silver Mill.
The settler tailing contains heavy unaltered ore which may be concentrated to recover heavy sulphides and particles of amalgam.
Costs. — ^The costs of pan-amalgamation with tank-settUng (Washoe process) in 1910 per ton of ore treated is:
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THE BOSS mOCBSS 243
I*bor 0.381
Loos of mETCury 0.7fiO
We&r of pans 0.200
Wear of dies and sboee 0. 400
Oil, interest, and luperinteadence 0 . 100
Tot»l coot per ton $2,363
One notes in p&rticxil&r tbe larger cost of supplies (cbemicalB, mercury, and castings) compared witb like items in gold milling.
THE BOSS PROCESS OF SILVER lOLLINO
This system, originated by M. P. Boss, a California engineer, differs from tbe Wasboe process in being continuous and generally requiring less labor. However, tbe AUis-Cbalmers Co. bas designed, for tbe Wasboe process, a wetr-cni^iiag mill in wbicb the settling-boxes bave sloping bottoms, so arranged that the content is transferred to tbe pans with but little labor. This takes away tbe advantage urged in favor of the Boss system. It may be added that the settling of the pulp in large tanks, combined with a mechanical system of excavating tbe content as in the cyanide process, ought to be efficient and labor-saving. "Hie Boss system may be applied to free milling ores and to refractory ores that need to be first roasted.
THE mOH-ORADE IHPlSSniG MUX, COBALT, OHTARIO
Tbe ore, containing natjve silver and argentite, together with the arsen-ides d cobalt and nickel (6 per cent Ni, 7 to 8 per cent Co, 40 per cent As), after being crushed to TO-mesb at the sampling mill, is deUvered at the plant witb an average content of 2800 os. silver per ton. It is fed to a tube-mill (see Fig. 149), 20 ft. long by 4 ft. diameter. Tbe charge consists of 3} tons of ore, 4^ tons of mercxiiy, and a 5 per cent cyanide solution. Tbe tube-mill is closed at both ends. Air, to accelerate chemical action, is introduced through a pipe.^ There is also an ingenious device whereby tbe excess of air is subsequently expelled. After nine hours in tbe tube-mill, 98 per cent of the sUver has been extracted from the ore, which, in tbe fonn of pulp, thei) passes to a settler, where tbe amalgam is separated by gravity. Tbence it goes to a deui-up pan and drainers. These last are canvas b^s for removing any excess of mercury.
The pulp and solution, deprived of amalgam, pass to a vat and are fed to a Butters filter, tbe clarified solution going to sine boxes where
1 As tbe result of the oxidation of the anmiides the temperature of the charge would rise to the bmUog-point were not the air supply to the barrel contr^ed.
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244 Amalgamation Of Silver Ores
the dissolved silver is precipitated on zinc sbavit^, thus obtaining an additional 2 per cent recovery. The shavings are in the form of coarse wire, necessary on account of the strength of the cyanide solution. The residue, left on the filter, containing 8 to 9 per cent cobalt, is afterwards sold for the value of this metal plus 85 per cent of the silver contents, so that from ore of 2600 oz. silver per ton, only about 4 oz. of sihner value is lost.
The amalgam, containing 80 per cent mercury and 20 per cent silver, is placed in retorts, each of which holds 450 lb. After the mereuiy has been distilled, the silver, still containing I per cent mercury, is taken to a reverberatory furnace. Here it is melted in a charge of 25,000 oz. After fifteen houra' exposure to a hot oxidizing atmosphere, without additicm of any flux, the molten metal is cast in ingots, each weighing 1100 ob. sUvet, which is 999 fine. Two oil-burners afford the necessary beat. The flue from the furnace is provided with a water-jet condenser, whereby 1000 to 2000 lb. mercury is arrested monthly. The gases escape at 100° F. During February, 1912, 550,000 oz. of silver was melted in this small plant.
The richness of the mine product under treatment and the completeness of the metallurgical operations leave a vivid impression. Within a small building it was possible to watch the successive sta^B by which a complex ore of a refractory type yielded its precious content in metal ttf such purity as to be ready for the mint. The entire process is so expeditious that the silver is delivered at New York within a week of the day when the ore is received at the mill and payment for the yield is received concurrently with the shipment. No less than 20 tons of mercury is in use at a given time. The cyanide has a cleansing action upon it; indeed, the xise of mercury would be impracticable without the cyanide, for the mercury would become " sick " or fouled, so as to hinder amal^- mation with the silver in the finely ground arsenical ore. The yoking of cyanidation and amalgamation constitutes another remarkable feature.
The right half of the flow-sheet. Fig. 149, gives a clear idea of the progress of operations. The product received consists of two-thirds hand-picked ore of 2800 oz. per ton, the rest, jig products of an average value of 2400 oz. per ton.
This is Tised on ores carrying silver, gold, and sulphides of tiie heavy metals, such as galena, blende, and pyrite, and sulphides which contain silver and gold. It is necessary that the silver not in the sulphides be amalgamable, as is silver chloride, argentite or native silver.
The process consists in wetr-stamping the ore, runnmg the pulp over apron-plates as in gold milling, concentrating the sulphides, which are
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Amalgamation And Concentration 245
shipped to the smelter, and, as in the Washoe process, pan-amalgamating the tailing and retorting the ama^am to recover the silver and gold.
Compared with either wet or dry silver-milling, the process has much to conunend it. The oie being lefractory, the wet process would recover little value. The tonnage stamped by the dry method with roasting would be low compared with wet-stamping, which is one and one-half to twice as rapid. It is tnie that by dry-«tamping and roasting we aie able to extract at least 10 per cent more metal than can be obtained by raw amalgamation, but thisis offset by the cost of treatment and the loss of precious metal in roastii^. The process also saves lead and removes galena, sulph-arsenides.
Tia. 139- — Stomp Mill Using Ama^natioii and Concentration.
and Hulph-antimotudes, all of which tend to foul and cause the loss of mercury. Such minerals are not amenable to amalgamation, and by removing them fpr smelting there results a cleaner or higher'^rade bullion. Manganese minerals that consume chemicals in the pan are also removed by concentration.
Amalgsmatkm and Concentiating Mill. — Fig. 139 is a perspective view of a 10-stamp mill. Let us suppose we are to treat an ore, in part oxidized, but containing the heavy minerals of lead and copper, with pyrite, arsenides, and manganese minerals. The ore contains the precious metals, a gangue trf quartz, calcite, and a little clay, and disseminated through it gold and the omalgamable ralver minerals cerargyrite, argentite, and native silver.
246 Amalgamation Of Silver Ores
Hie purpose is to save the piecioua metala by {date and pan-amalgBnuitian, and the heavy minerals with silver and gold by coQcoitration. Scane of the silver and gold escapee recovery and is lost in the tailing. Since 8iilid>- arsenidee and manganeee minerals are moeUy ronoved, they do not interfere with subsequent pan-amalgamation where arsenic would sicken the. mercury and manganese consume chemicals.
Tlbe ore and water are fed aut(»natically to a lO-stamp battery, eadi stamp crushii^ 4 taas per twenty-four hours to paee a SO-mesh screen. The pulp issuing from tiie mortar Sows over two apron-plates (one for each five-stamp mortar) and a part of the gold and silver is recovered. Ttie flow is distributed evenly to four concentrating tables at a lower level, the concentrate (10 per cent c£ the whole) being separated to ship to smelting works, while the tailing is carried to the ten settling-boxes in a double row. These are seen at the left ot the pans. The distribution is into a double launder between the two rows. By drawing the plugs in the bottom of tbe launder, the flow can be directed into any box demred. From Hob point on, tbe operation is ctmducted as described for the Washoe prooees. There are four amalgamating-pans and two settlers. Bhiestone and salt are used to dectHnpoee tbe argentite. Mercury or amalgam escaping tbe apron-i^atee finds its way into tbe settling-boxes and thence to the pans, and is more thoroughly recovered than if it depended upon obtuning it in the concentrate as in gold-milling. The four 5-ft. txxnbination grinding and amalgamation-pans each treat 3000 lb. per chai^, and with a foiu'-hour treatment, this equals 36 tons daily, which with ti» 4 tons of concentrate already mentioned is a 40-ton output of tbe mill. Some ores, not so readily treated, take six to eight hours, and lessen the capacity of tbe mill accordingly.
In a certain ore of this kind, containing 0.40 os. Au and 9.02 Ag per ton tbe recoveries on the apron plates were 22 per cent of the gcJd and 3 per cent (tf tbe alver respectively; at the concentrating taUes 28 and 32 per cent; in the pans and 32 and 35 per cent; loet in the tailing 18 and 30 per cent. Of the lead and copper 85 per cent was saved in the ccmcentrate.
Ctmcentrsting adds but little to the cost of this milling, so Out $3 per ton may be taken as a fair estimate in 1910.
The Chloridizino Roasting 09 Silver Orbs
Silver ore contuning sulph-arsenides, sulpb-antimonides, or tetrahedrite, cannot be treated directly by amalgamation nor by cyaniding. Such ore is subjected to a roast with salt to convert the Edlver into a chloride, before it can be successfully treated by these methods. The above minerals are often accompanied by pyrite, blende, chalcopyrite, and galraia.
Preliminary to roasting, suclk oie is dry-crushed, either hy ndls or by
CHLORIDIZING ROASTtNQ 247
stamps. Ores contamiug galena and blende are preferably crushed to 40-meshBue, thoseh&vingpyritetoSto lOmeeb. The " Roasting " ia done in a reverberatory furnace, and requires the use of atdt. There must alBo be 3 to S per cent pyrite present to furnish euli^ur for the reaction, and if the ore does not ctrnttiin this, it must be added. If more than 8 per cent sulphur is present, the percentage is reduced to that point by roasting before the salt is added. The amount of salt required varies according to tiw quantity aS copper and iron sulphides present which consume the evolving chlorine.
Chlcffidizing RoAStlog. — This operation is at first an oxidizing one conducted at the temperatures specified in the chapter on roasting. The action ia chiefiy upon the heavy metals, converting them into either (mdes or sulphates. It may be divided into three stages: (1) the kindling, (2) the deeulphurization, and (3) the chlorination of the ore.
Hrst Staga. — In the first or kindling stage we find the loosely held sulphur being driven off, and the ore taking fire, producing a blue flame.
Second Stage. — In the second stage, the air oxidizes the sulphides, and particularly the newly formed iron sulphide. Reacting upon the sulphantimonides and arsenides, it volatihses them and removes them from the ore. Copper and iron sulphates are also formed, the latter according to the following reaction:
Third St>ge. — In the third stage, at 590° C, the sulphate formed in conjunction with air reacts upcm the salt, thus:
The chlorine thus liberated acts at once upon the silver compounds and converts them into chlorides. Thus:
Zinc blende becomes oxide and zinc sulphate, while sulphur dioxide escapes. Galena and sine sulphate remain inactive and fail to decompose the salt. They roast slowly, while pyrite, in presence of salt, decomposes quick^, and generates chlorine at a period in the roasting when neither the blende nor galena is sufficiently oxidized to expose silver to the action of the chlorine. If, therefore, the salt is mixed with the ore at the battery, the chlorine generated by the reaction of the ferrous sulphate and salt is lost, an imperfect chlorination results, no matter how loi^ roasting is continued, ntM* how much salt is added. Hence in roasting an ore containing
248 Amalgamation Op Silver Ores
blende and galena it ia of the greatest importance to add the salt later and not at the battery. On the other hand, if the roasting continues until the sulphides are well oxidised, the iron sulphate decomposes and no chlorine is generated, and again we have a badly chloridized ore. The desirable time to add the salt is after continued roasting at a low heat that does not break up the iron sulphate. This is shown when the black color of the ore changes to brown, but shows still the presence of black particles. A distinct odor of chlorine is then to be noticed, due to the decomposition of the salt. The best results could be obtained by adding a mixture of green vitriol (ferrous sulphate) and salt; but the ore would hardly justify the expense..
The salt is added to the dry ore at the time of charging, if the percentage of sulphm is suitable, or later if the excess of sulphur must be first removed by roasting. The temperature is increased only gradually to kindle or start the ore to burning and to begin oxidation. As the temperature rises oxidation and the formation of sulphates occur, and at the necessary hi^ temperature these act upon and decompose the salt and chloridize the ore.
Heap Chlorination. — It is not considered necessary to continue the roasting to convert all possible ralver into chloride, but to withdraw the charge while hot before this st^e is reached. During the gradual cooling (twelve to thirty hours) further chloridizing proceeds, due to the mass action of the free chlorine, with which the ore is saturated, acting on the undecompoeed silver sulphide. This may increase the cbloridizatiou 10 to 40 per cent.
Upon completion of the operation of " heap chlorination," as it is called, and with ores containing copper chloride, a wetting down or sprinkling causes an additional chlorination of 3 to 6 per cent. Thus at the Lexington mill, Butte, Mont,, the ore, after roasting in a Stetefeldt furnace, was chloridized to 65 per cent, after two hours in the heap to 75 or 80 per cent, and at the end of thirty-six hours to 92 per cent of the silver content.
The loss in silver by volatihsation, when the ore has been property and carefully roasted, should not exceed 8 per cent except in presence ol volatile elements like arsenic, antimony, selenium, or tellurium. If, bow-ever, the roasting is completed at a high temperature the loss may rise to 18 per cent.
Remarks on the Chloridizing Roast. — The most difficult, and at the same time the most important process for the treatment of base mlver ores by wet methods, is undoubtedly chloridizing roasting. It is always the safest plan for the operator to roast as thoroughly as possible. If the an is well chloridized, sodium hyposulphite or cyanide extracts all the silver chloride. A high chloridization does not necessarily involve a high loss I^ volatilization. It is well suited to refractory manganese silver ores, as
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Dry Silver Milling 249
most of the silver b converted into a readily soluble silver chloride. It has never been favored for gold ores on account of high volatilization losBes; in fact, this is also the weakest point with «lver, since silver chlorid is quite volatile.
Chloiidizing b; Bltst-^tMSting. — The Dwight-Lloyd machine, Fig. 83, bids fair to be successfully used for a chloridizing roast. It is claimed that the volatilization of the silver is entirely under control, and moreover the cost of roasting is low.
This process for the treatment of rebellious silver ores, in which the metal is so locked up as to require roasting before it can be amalgamated, was developed at Reese River, near the Comstock Lode at Virginia City, Nev. The ore contains silver sulphide, particidarly the antimonial sulphides, and the sulphide of the base metals such as copper, iron, zinc, and lead. Galena, however, if present exceeding 5 to 10 per cent, renders the ore unsuitable for chloridization.
The treatment in brief consists in dry-crushing and roasting the ore then amalgamating in pans to recover the silver and gold. The drycruslui^ is done either with rolls or stamps. Crushing with rolls is described in the chapter on Crushing. If dry-stamping is employed the work is done in the dry-crushing silver mill.
The patio, or Mexican amalgamation process, was introduced by Medina into Mexico as early as 1557, and has been practiced in that country down to the present time. The ores best suited to it are silicious ones carrying finely disseminated native silver, silver sulphide, and chloride. A limited amount of pynte, galena, cerussite, or the copper minerals may be present without serious interference with the process, but much blentte causes low extraction. Where any gold occurs this is not recovered.
In outline, the process conrasts in finely crushing the wet ore, and treating the mud or fine product in a flat pile in a large paved yard or patio, Bait and bluestone being added upon the pile and well trodden in by mules. Mercury is next sprinkled on and mixed in the same way. The above operations require two to four weeks. The fine product is then washed in tanks to separate the heavy mercury and amalgam from the light tailing, and the amalgam is recovered and treated as in silver milling.
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CHAPTER XXII SLVrai HILLING BT HYDROMETALLURGICAL PROCESSES
PBHTCIPLBS of the HTDROMBTALI.niLOT OF SILVER
A wet-procees for the recovery from the ore consists in dissolving the metal by means of a solvent and precipitatii^ from the solution in a can-- venient form. The silver compounds which can be obtained readily in solution are the sulphate and the chloride. In cyanide solution argentite is readily soluble, while luby silver, freislebenite, and stephanite, are sparingly bo, though readily soluble in mercurous potassic cyanide. Silver sulphate is soluble in hot water, while silver potassic chloride is dissolved by brine solution or by sodium hyposulphite (thioeulphate). From the aqueous solution of the sulphate Eolver is precipitated by metallic c(q>per; from the brine solutitm of its chloride by copper, or when in dilute sohiticm by zinc, iodide; from the hyposulphite solution by sodium sulphide; and from the cyanide solution by metallic sine.
The AugUBtJn and the Ziervogel processes, introduced in 1840 to 1850, were used in a limited way almost exclusively for the treatment of matte. However, the recent apphcation of the Augustin process in connection with blast roasting for the treatment of low-^rade complex silver ores contiuning lead and copper is worthy of note.
The Augustdt Pkocbss
This has been used for the extraction of silver from ore and from copper-bearing matte, obtained as a product of smelting. At Koeaka, Japan, ore consisting of one-half heavy spar and containii^ 10.5 os. silver per ton is thus treated. The ore is crushed and roasted with salt in a reverberatory furnace, and, after drawing from the furnace and moistenii^ on the cooling-floor, contains 80 per cent of the silver in the form of chloride. It is leached with a hot 18 per cent salt solution in regular leaching-vata. The leaching is continued until a polished plate of copper shows no precipitate of silver when held in the flowing filtrate. It requires 0.66 ten (rf brine to leach a ten of the ore. The sand is washed with hot water, and the tailing rejected.
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The Ziervogel Process
Thb Ziervogel Process
This procesB, practiced at Mansfeldt, Geimany, and at the BoBtou & Colorado enieltiiig works at Ai^, Colo., is adapted to the treatment of rich copper matte containing little or no arsenic, antimcmy, or bismuth, any of which would form insoluble compounds with silver. The method may be divided into three parts: the roasting for silver sulphate, the leaching, and the precipitation of the silver.
The ProceBS. — Referrii^ to the flow-sheet of the process (see Fig. 140)
Fia. 140. — Flon-flheet of Ziervogel Pmcem.
we have in furnaces A , the operation of producing the matte or legulus frcHU golf^ and silveF-beariug copper ores. The details of the proceas are described in the chapter on the Metallurgy of Copper, under the head oS " Reverberatory Matte Smelting." The ctHnpoettion of the matte is Cu, 47.3 per cent; Pb, 8.1; Zn, 2.7; Fe, 17.7; S, 21.6 with 400 oz. silver and 15 OS. gold per ton.
Preparation of tiie Matte. — The matte is crushed and passed through rolls at B to reduce it to &4Qe8h size, and sent to a reverberatory furnace C, where it receives preliminary roasting. The roasting reduces the sulphur to 6.3 per cent, and ctmverta the iron and copper sulphides to the corre- Bpcaubng oxides and sulphates, as described in tiie chapter on the chemistry
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252 Silver Milung By Hydrometallurgical Processes
of Oxidizii^ Koastii^. This partly roasted product then goes to a ChiUaji mill D (see also Pig. 46), where it is finely ground to 60-mesh.
Sulphatizmg Roesting. — The partly roasted iaatt« is next treated by hand in charges <d 1600 lb. by a sulphatising roast in small single-hearth reverberatory roasters at E. In the process the iron and copper remaining in the form of sulphides are converted into sulphates which react on the silver sulphide at a slightly higher temperature, as follows:
(10) AgaS+30+CuS04 = Ag2SO*+CuO+SOa.
It has been found that the addition of 2 per cent sodium sulphate (salt cake) faciUtates the change. The roasting takes place in four stages as shown below.
During the first stage, of \\ hours, the draft is checked, the side dooFS kept open, and the charge held at a low temperature. The charge becomes evenly heated throughout, and glows from the oxidation of CU2S to CU2O.
During the second stage, of 1^ hours, the heat is increased and the charge constantly rabbled. Iron sulphate is decomposed with the consequent formation of copper sulphate. The chai^ swells and beccones spongy by the formation of this salt.
In the third stage, the temperature is increased for an hour until tests show that the silver is " out," that is, in the form of sulphate. The following reaction occurs :
During the fourth stage the temperature is kept constant. The charge is gathered and pressed down with a heavy, loi^-handled iron paddle to break the lumps, and then vigorously stirred to oxidize the remaining CuaO to CuO, and decompose copper sulphate. The temperature is not further increased, since it would decompose silver sulphate, forming silver oxide, rendering the silver again insoluble.
The progress of the roast is tested by dropping small samples from time to time into hot water. Soluble sulphates dissolve in the hot water; and in the tests made eaHy the solution becomes deep blue. Later, as the silver sulphate betpns to form, it is immediately reduced to alver spangles by the cuprous oxide present. As the roasting advances during this stage, the copper sulphate decomposes, and the sohition becomes less bhie in the test and the silver spangles increase and aftetward diminish. During the fourth stage the CusO is changed to CuO and the spangles no longer show. A lightrblue color of the solution remains, due to the presence of a httle copper sulphate, which indicates that the silver sulphate is not itself becoming decompoeed. A sample thus roasted showed by analysiB
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The Ziervogel Process 553
^-5 per oent FeSO* and ZngO*; 0.6 per cent CUSO4, and 1.73 per cent ^j^^O* C348 08. silver per ton), bo that there was left in the matte (there ■^ii^g tio loss of weight in roasting matte) 52 oi. per ton or 13 per cent of the silver in msoluble form.
■^-^aclaing. — The roasted matte is charged into tanks F and leached
^'^*" hot. water to dissolve the sulphate above described. The filtrate goes
. a. series of boxes H, contaiiung copper plates upon which the silver pre-capita
-tes m the fonn of white shining crystals. The alver-free solution,
coTi-taij^jjjg in addition to the original copper sulphate that which it has
■t-aken from the copper plates, goes to tanks /, where the copper is precipitate
^i -upon scrap-iron to recover the copper. The final solution is rejected.
THe cement-silver from the precipitating boxes is transferred to a tank,
a>n<i dilute sulphuric acid is added. It is boiled by forcing in a mixture of
*-ir axid steam from an injector. The treatment oridizes and dissolves the
*^"**=^s ot copper still retained by the silver crystAfa, and keeps them in
^K^'ta.tion at the boiling temperature of the acid nurture. The copper sul-
■pnate solution is now run off and the residue repeatedly washed by decan-
■*-a.t.ion with hot water to free it entirely from copper. It is transferred to a
»0¥\g pan over a coal fire for drying and is then jnehed down in crucibles
«i a wind-furnace and obtained in ingots 999 to m^ fine.
Residue from the Leaching Tanks. — The eifracled residue remaining
wi the tanks F, still retaining 52 oz. silver per ton »s above stated, freed
'■*'ni sulphates, and composed mainly of iron and rapper oxides, is sent to a
verberatoiy furnace K, to form copper raalte. TV slag produced in iht
I * "^^"tBoesbacktotheore-smeltingfunufe.^.while the matte, tappH
* I ,?®***^ molds, is sent to the reverberatoiy knaer L, to be treated !;t
laj
"SUsh process of making " best-seleded topper." Here the mATif.
J through the fire »^r ttmace. TTw e?*^*; »■ the drops *>f!:^,:->--^^^ wbi^le ch:ire*' c*— -^ acts on thre '.^l^ *=-'''
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264 Silver Milung By Hydrometallurgical Processes
Bome of the silver. On the other hand, the Bupenmtatit matte has risen to the grade of wbite^netal of 75 per cent copper, and carries 90 to 100 ok. silver but not more than 0.2 oit. gold per ton.
To prepare it for the extraction of the silver, the matte is again given a Bulphatizing rooBt, but in a different furnace from the one used for the first matte. The residue after this second treatment, principally a copper oxide containing 10 oz. silver per ton, is sold to the oil refiners. The bottoms, formerly treated at the Argo works by a secret process for the extraction of the precious metals, is now electrolytically refined.
THE HTPOSTTLPHITE LIZIVIATIOn OF SILVER ORB (VOIT PATERA . PROCESS)
Hyposulphite lixiviation can be practiced upon ore containing simple or compound sulphides of silver that have undergone a preliminary chloridizing roasting. The silver sulphides, in the roasting, become converted into silver chloride. The process also applies to silver ore already containing the silver as chloride. Free-milling ore, such as oxidized ore containing the silver in the native state, or as chloride, or to some extent as ai^ntite, are preferably treated by miUing and amalgamation. Native silver and silver sulphide in a favorable form can be recovered by milling uid amalgamation, whereas by hyposulphite extraction, they would remain insoluble. The process is not suited to the treatment of gold ore. The extraction of gold is low; usually less than 50 to 70 per cent. One of the most useful applications is to the treatment of argentiferous blende that has been hand-picked or concentrated from galena, and may still contain lead up to 8 per cent.
The hyposulphite process is based upon the fact that silver chloride readily dissolves in dilute solutions of sodium hyposulphite. The chloridizing roasting is imquestionably the most important part of the process, and the chief attention and study is to be given it.
It consiste in crushing the ore, roasting it, and treating the roasted ore in filter-bottom vats, first with water to remove the soluble chlorides and sulphates of the heavy metals (base metal leaching), then leaching with a dilute solution (rf sodium hyposulphite to dissolve the mlver chloride. Silver sulphide is precipitated from the filtrate with soditmi sulphide, dried, and roasted to remove the sulphur, and the residue is sent to the smelting works, or treated in an Ilnglish cupelling furnace.
This is a modification of the Patera process, principally by the use of another solution, in addition to the hypo-solution, for the extraction of the silver. By mixing in solutioQ two parts of the hypo-salt with one of copper
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The Russell Process 265
sulphate we obtain a double salt NaaSaOa ■ CuaOs, called the extrarflolution. It has a solvent power nine times as great as that of the ordinary hypo-solution for native silver, diver sulphides, silver arsenides, and silver antimonides. In the case of an imperfectly roasted oie the use of the extra-solution insures the extraction of more silver from the compounds meotioned tiutn could be obtained by the use of ordinary hypo-solution.
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Chapter Xxiii Cyanidation Of Silvbr Ores
PRINCIPLBS OF CTAniDATION
Silver ores carrying gold, and in which the silver occurs as chloride, or ai^entite, or stepbanite, have been successfully cyanided. Manganese silver ores contain the silver intimately associated with the manganese and will give no extraction when treated raw, unless there be a preliminary acid treatment.
Of the silver minerals, native silver, in particles so large as to be viable, is insoluble in potassium cyanide in any reasonable time. Silver chloride, bromide, and argentite, are readily soluble. Ruby silver, etephanite, and freieslebenite are sparingly soluble in potassium cyanide, but readily soluble in mercurous potassium cyanide solution.
A chloridising roast is always beneficial to ores containing silver sulphides as it increases the percentage of extraction. However, it is not essential and is seldom employed in the cyanide treatment of silver ores.
In handling silver ores the reactions are more complex than in treating ores of gold, owing to the greater chemical activity of the silver compounds, and to the fact that owing to the lai^r quantity of contained metal the cyanide solutions are necessarily stronger.
Important matters in cyaniding silver ore are the following:
(a) A long time, ten to twenty-five days in the case of sand treatment, ' is needed for leaching. For slime treatment from forty-eight to ninety-six hours would suffice for a complete cycle, in which time a h^her percentage of extraction would be obtained than by a fourteen-day treatment of the corresponding sand. The silver compounds are not so easily soluble as gold, and a larger amount must be dissolved.
(b) Thoroi^h oxygenation is necessary, not only because of the large amount of silver present, but because silver compoimds need at least initial oxidation to become properly soluble in cyanide sohition. Hence an advantage is secured by the double treatment of sand. Also during leaching, if the solution be allowed to sink several inches below the top of the chai^, before another wash-solution is run on, air is drawn in and penetrates the ore, and the solution following forces the air downward through the ore. In the treatment of the slime the pulp may receive thorough aeration by agitating with air.
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Precipitation Of The Silver 257
(c) Stronger BolutioD is \ised than for the treatment of gold ore. Thus the first or strong solution may be 0.7 per cent, the weak one 0,26 per cent, ^vhile for gold ore, a 0.5 per cent eolution would be called strong and 0.05 per cent weak.
(d) The consumption of potassium cyanide is higher than in the treatment of gold ores. It varies from 1.5 to 4 lb. per ton as compared vrith 0.4 to 0.8 lb. consumed in the treatment of gold ores.
(e) The precipitation of silver from cyanide solution by sine shaving presents no difficulties and is practically complete. Despite the fact that a relatively great amoimt of silver has to be precipitated, as compared with gpld, no more zinc is consumed.
The Precipitation Of Silver Proh Cyahide Solutioit
Silver in cyanide soluti<m may be recovered by precipitating upon zinc shaving or by zinc or aluminum dust according to the reactions (6) and (8), page 147. In presence of copper the precautions described imder head of " Precipitation of Gold from Cyanide Solutions " will equally apply to silver precipitation. Silver ores of course make a much larger bulk of precipitate than gold. On this account the Merrill precipitation process is preferred to the uae of the zinc boxes illustrated on page 258, though for a small plant they might be used- Fig. 141 shows in outline the course of procedure for the clean-up, the pressing, drying, and melting of the precipitate from clarified silver solutions. The novel feature is the washing of the fine precipitate from the HOC box with pregnant solution into the filter press, where it is collected. The pumping of pregnant solution through the filter press is continued at intervals, until the tail solution ris^ to approximately the value of the pregnant solution precipitated. During the earlier stages of pumping, the tail solution from the filter press is of low enough grade to divert to the milling solution, but, during the later stages, it is advisable to return it to the head of the zinc boxes. With solutions in which silver occurs in the ratio of 50 parts to 1 part of gold, precipitate is at times obtained which contains approximately 90 per cent of silver and gold. There is no difficulty in obtaining precipitate, regularly, well over 80 per cent of silver and gold. When of this grade it is very eafdly melted, with a minimum consumption of fluxes and fuel, and, of course, produces a h^h- grade bullion. At a certain Mexican plant, where this plan was adopted, the grade of the precipitate was very materially raised and the fineness of the bullion was increased from 800 to 900. This resulted in a monetary saving of several hundred dollars per month. A similar plan appears to be feasible with zinc dust precipitation. It is still a question as to how far such a phn could be carried with solutions in which gold predominates.
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CYANlDATIpN OF SILVER ORES.
With the Merrill proceSB double precipitation ia generally practiced, as shown in Fig. 1 17 ; the apparatus is in duplicate, constituting two circuits.
ji'
i
Is
i
L
1
11
¥
1
In the one circuit the heavier feed of zinc-dust is maintained to ^ve a barren solution, while in the other there is just sufficient to precipitate the bulk (rf the silver. The amount of the barren solution to be made is
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The Santa Gertrudis Refinery 250
detenniued by the needs of the final washes, while the other, needing less zinc-dust and carrying a few cento worth of silver, is used in the agitation tanks. Thus in a Mexican silver-mill <A 500 tons capacity, about half the solution was completely precipitated to retain but 1 to 2 cents worth of silver per ton, using 1.1 02. of sine-dust to one of silver. On the other hand the partially or almost preci)Ntated solution needed but 0.74 os. of dust to one of silver and this retained but 10 cents of solution, well suited for re-xise.
THE SAITTA OBRTRCDIS PKBCIPITATIHG AUD KBFIlflHO PLANT
Fig. 142 is a plan of this installation, showii^ the method wed for double precipitation by the Merrill process.
There are four pregnant solution storage tanks, so that an exactly measured quantity of solution can be precipitated by a known weight of zino-dust. Owing to the large quantity of idnc-dust needed this is slowly fed to the zinc-dust mixing cone by an endless belt throu^ a worm^ear drive, this belt taking its feed from a supply-hopper in regulated quantity. Referrii^ now to the partial preciintaticai circuit, the two tank-discharges unite in one just before the zinc-feeder. The emulsion from the mixing cone drops by a vertical pipe into the tank discharge which constitutes the 12-in. suction of pumps Nos. 1 and 2 of the circuit. From the pumps the solution now impregnated with the zincdvist, passes to the Merrill piesaes Noe. 1, 2, 3, and 4. When the desired tank, Buch as tank No. 9, is emptied and the solution, now nearly freed from its Eolver, is filtered and washed, the filtrate goes by a 10~in. pipe line to mill-solution sump-tank No. 16. The emptied presses are blown with air to drop off the precipitate on the filter leaves and the press opened for the removal of the precipitate. For the production of barren solution a similar procedure is followed, taking the pregnant solution from tanks Nos. 1 1 and 12 an<'. delivering through presses 5, 6, and 7 to the barren solution sump-tank Vo. 14.
Fig. 115 is a viewof the styleofpressused. The filter press frames are triangular, the pulp entering at the bottom <^ the frame. The solids gradually accumulate on the filter cloths of the frames, providing a imiform filtering layer of a fine-drained precipitant throu^ which the solution must pass to ensure contact between the zinc and the gold particles for instant precipitation. The barren or nearly barren solution, the product of the press, then passes by way of the discharge launders to the completed precipitate or to the partially precipitated solution tank.
For a clean-up the press-cakes are dried by forcing air through the precipitate on the filters, the press is opened and the caked niaterial dumped into tram-cars. The product is sampled, weighed and determined for contained moisture, thus giving its exact dry weight.
Cyanidation Of Silver Ores
We give in Pigs. 116 and 1 17 views of a press-frame and trf a press-plate. There are say forty such frames in s press. The plate is covered on both of il« faces by a filter cloth. When the f runes and plates are placed ti^tJy
together in the press the common circular passage along the top admits the solution to each frame It is led to the bottom of the fruue by the drop pipe there shown Pa.s8ing through the filter cloth it escapes by a cock at the left-hand upper comer of the plate.
Refining Silver Precipitato
Precipitation By Alumiituh Dust
This IB not new as far as the actual knowledge of its use ie concerned ; but it was only recently that its use in a practical way on a large scale was studied. This is done at Deloro plant and the Nipissing mines, Ontario, where rich silver ore containing arsenic and cobalt is treated. They are perhaps exceptional cases, rendering its use necessary.
Precq>itation at Nipissing. — It was found here that, after solutions had been precipitated by sine they rapidly lost their dissolving efficiency, so experiments led to the use of aluminum dust. The metal must be well agitated with solutions before precipitation. The fact that aluminum does not replace the precious metals in the cyanogen compound renders necessary the presence of a caustic alkali. The reaction in precipitation is probably
(13) 6NaAgCN2+6NaOH+Al = 6Ag+12NaCN+2Al(OH)3.
From this equation it will be seen that there is a regeneration of cyanide.
At this plant the pregnant solution is pumped to a sand-<er. It is then precipitated by the Merrill process as above described. Some 550 to 600 tons of solution are handled daily, the head-assay running about 8.25 oz. the tail assay or barren solution 0.10 oz. of silver per ton. Precipitation averages 98 per cent, i^ng 0.02 lb. of aluminxmi per ounce of silver at a cost of 18.5 cents per ton of ore treated.
DRYmO AND REFTRIRG SILVER PRECIPITATE
The early practice of selling the precipitate to smelters is still adhered to by a few companies, but the majority convert it into bullion before marketing. Cyanide bullion varies greatly in fineness, depending upon the character of the ore treated as well as the equipment provided for refining and the degree of skill exercised in its use. It is, obviously, not good practice to carry local refining to the point where the cost is greater than the advantages to be realized from marketing higher-grade bullion. The only case that I know of where fine bullion is produced which requires no further refining is that of the Nipissing Mining Company, Cobalt, Ont., Canada. Here, imusual conditions make the production of fine bullion a comparatively easy matter.
Acid Treatment — Prelbninary acid treatment of the precipitate has been found unnecessary in most cases where silver predominates, but is still adhered to in the majority of cases for gold precipitate.
The Tavener Hethod. — The Tavener method of refining, involving the melting oi the precipitate in a small reverberatory furnace with various fiuxes and lead, followed by cupellation. is still in general use in South Africa, but has not gained ground in this country.
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Cyanidation Of Silver Ores
Ct^elliDg. — ^The Homestake method of refiniixg, involving acid treatment and briquetting with lead flux, followed by melting and cupellatim in the ordinary English cupel fmnace, ae well as tite treating of all the byproducts in a small blaat funiaoe, is used by a few American plants.
BU8t4iinuu;e Treatment — ^The melting of tbe precipitate, briquetted with iim proper fluxea, in a small blast-funiace, fc^owed by cupellation of the lead, has been found of advantage when dealing with a large v<dunQe of low-grade precipitate. Loesefl from dusting, which m^t appear to be the chief objection to this method of melting, are claimed to be insignificant when a fnoper flux aystran is provided. This method of melting is more econ(Hmcal thim either ihe Tavener or HtHnestake practice.
The Electric Furnace.— The electric furnace has also been used in a few cases.
The Tilting Furnace. — The most simple and satisfactory method of converting the high-grade precipitate which is obtained, by proper manipulation, from ores in which silver predominates is to melt the precipitate directly, with the minimimi proportitm of flux, in the tilting type of furnace. A few mills use a double-chamber tilting furnace, in which tbe flame comes in direct contact with the charge being melted. This furnace is perhaps more economical of fuel, but, ualess tbe precipitate is briquetted there is greater risk of loss through dusting than when the precipitate is melted in a closed crucible. In the case of silver precipitate, with proper manipulation in the crucible furnace, it is a serious question whether it pays to briquette, the losses being less under these conditions than the cost of briquetting.
Let us take the case of a low-^rade precipitate contaming 25 per cent gold and silver; 39 per cent lead, 20 per cent zinc, \ per cent copper; \ per cent sulphur, 3 per cent lime and 1.4 per cent silica. In the following table we will have:
s
On
a«.
Wawbt, PouDda.
Combinxl u.
And, Pound..
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Refining Silver Precipitate 263
Here the ratio of acid to basic oxygen is as 2 to 1. There was yielded 134 lb. of sl^ and 5^ lb. of a white-looking layer or cover of sulphate of soda. Of niter there was used 35 lb., of which 6 lb. was needed for the cover, leaving 29 lb. to give the 22.5 lb. of soda for the slag. Of bonuc 39 lb. was required to yield 27 lb. of BaOa, so that there was a total of 93 lb. of fluxes for 100 lb. of precipitates. When melted the resutttoit bullion was 972 fine. A plumbago crucible is preferred, but where there is much copper there is a clay crucible used because a plumbago crucible would tend to reduce the copper which would then enter the buUion, reducing its fineness. Types of Furnaces Uaed.— Various types of furnaces are used in melting precipitate, namely, the ordinary wind-furnace holding from a 60-to 200- Hize crucible (see Fig. 1) ; the Faber-du Faxir tilting furnace as used at
Fio. 146. ^Monarch Rockwdl Refining Furnace.
Ka^oorlie; the Steele-Harvey oil-fired tilting furnace; the Monarch- Rockwell as used at the Belmont, Tonopah, the latter as shown in Fig. 146. Drying, Meltmg and Refining by Fusion. — ^The damp precipitate from the ore is well mixed with a calculated amount of fiiixes, the moisture loaded into pans and placed in a large muffle furnace and gradually heated to a red heat. Moisture is driven off and the precipitate in the acid settles to about one-third of its original bulk. The sintered mass, free from dust, gives an excellent product for crucible melting. The fluxes used are generally Chili niter (sodium nitrate) for oxidizing soda-ash to take up the sulphur, borax, and sand as acid fluxes for the bases. These are added in such proportions as to produce with the silica and the bases in the precipitate a slag of two of acid to one of base oxygen. The ratio
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264 Cyanidation Of Silver Ores
of borax .glass to silica sboiild be as 2 to 1 ; or in presence of much lead aa Itol.
Smelting in on OU-burning Reveiberatory Furnace. — The hearth of the furnace is 1 1 ft. by 4 ft. 9 in. wide and will take a cbai^ of 5 tons of pi«- cifntate. Before using, the furnace is seasoned by melting in 700 lb. of old slag which consolidates the bottom. The precipitate is property fluxed and melted down, giving about 2) tons of bullion of 883 fine in ralver. The furnace is operated as often as a 5-ton charge is ready, the smelting taking from sixteen to twenty-eight hours.
Chbhistrt Of The Ctaitide Process For Silver Ores
When a solution of potassium cyanide is brought in contact with finely divided silver or silver chloride the metal is dissolved according to the following equation:
The reaction is similar when the cheaper sodium sulphide is used.
Oxygen is needed to complete the reaction, ao that the ore-pulp should be properly Ei6rated. This may take many hours. When the occhided oxygen is used up the reaction ceases, but resumes with a fresh supply of air. As described for gold ferrous and ferric sulphates result from the decomposition of pyrites and tend to precipitate silver. To overcome this an addition of quicklime is made m more than sufficient quantity to overcome the acidity, the excess being termed " protective alkalinity." When silver-bearing sodium cyanide solution is brought in contact with zinc shavings or 2dnc dust we have the reaction :
(15) NaAg(CN)2+2NaCN+Zn+H20 = NaZn(CN)4+Ag+H+NaOH.
One part of zinc is calculated to precipitate 1.7 parts of silver. But cyanide is also used up by direct combination with the zinc-
In both these reactions there is an escape of hydrogen bubbles. When aluminum is used as precipitant we have :
(17)2NaAg(CN)3+4NaOH+2Al=4NaCN+2Ag-hNa3Al204+4H,,
in which one part of aluminum precipitates four parts of silver, in this respect being so much more efficient than zinc'
1 In place of writing sodiuin cyanide NaCN it is often written NaCy, the Cy being a oonvmient way of specifying the CN.
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The Belmont Mill 265
TmCAL SaVEIL HILLS
The following are descriptions of silver mills:
BELHONT HILLmO CO. HILL, TONOPAH, NEVADA
The ore is a fine granular quartz of about 72 per cent silica with a few per cent of sulphides. The silver occurs as diaaeminated silver sulphides, antimonides, and selinides, a little native silver, and a little gold. Thus to 0.32 oz. of gold there would be 32 oz. silver, or in weight 1 of gold to 100 of silver. The ore is treated by concentration and the tailings agitated in cyanide solution and filtered. The operations may be divided into five; (1) picking, coarse crushing, and dehvery to the mill bins; (2) stamping, tube-milling, and concentrating to remove the heavy part to be sold to the smelter; (3) cyanlding the tailings from the concentrating tables to yield a pregnant solution; (4) precipitation of the gold from the solution; (5) Defining the precipitate.
(1) Pickiag, Coarse Crushing and Conveying. — ^Thisis done in a separate building. The ore, broken underground to 9-iii. size or lees, is stored in two 1000-ton fla1>-bottom bins. It forms a natural slope, so about half of it can be drawn off at the side gates through shaking grizzlies, having 2-in. openings. The undersize goes by a conveying belt 6, to a trommel 8, while the oversize falls upon an endless picking belt, 40 in. wide, traveling 45 ft. per miniite. Here the ore-sorters pick out the waste laid throws it down chut^ to a 20-in. conveyor belt, which discharges upon the waste dmnp. The picking belt discharges upon a shaking feeder which feeds it to a gyratory crusher 8. The crusher discharge joins the imdersize from the shaking grizzly in the trommel, 4 ft. diameter by 14 ft. lot^ that has l^-in. openings. Its imdersize goes direct to the inclined conveyor belt 11, and the overaze feeds the gyratory crushers set to crush 1-in. size. The discharge from the crushers joins the trommel undersize on the inclined belt-elevator.
Sampling. — An inexpensive sampling of the ore is thus effected. On a vertical shaft is a bevel gear with a bucket of 5^ lb. capacity attached to it. When in the revolution of the gear, the bucket traverses the stream of falhng ore from the top end of the inclined conveyor it takes out a sample. Now, on the gear, one-third of the teeth are removed and at that space a counter-balance actuates the gear, allowing the bucket to make a quick cut through the ore. In this way a sample of 1 ton in 500 is taken to be cut down by the usual sampling methods.
(2) Stamping, Tube-miQing and Concentrating. — ^From the head of the inclined belt the ore is conveyed by a horizontal belt and distributed evenly to flat-bottom mill bins 16 ft. wide, 17 ft. high, by 110 ft. long, of 1500 tons capacity. Half of the ore which will run freely is fed to the stamp-batteries.
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Cyanidation Of Silver Ores
rreih Water Tuiki
Fia. 147.~Flon-«heet at Belmont Mill.
The Belmont Mill 267
Cmshmg is done by sixty stamps of neariy 9 tons duty per stamp through 4-to 6-mesh screens. It is done in cyanide solution at the rate of 5 toDB and with the addition of 1 lb. quicklime per ton of ore. The battery is ^own in Fig. 87.
The stamp battery discharge is detirered to eight Dorr duplex-classifiers placed in closed circuit with their respective tube-mills as ^own in F^. 40. A further addition of lime as milk of hme is made at this point, also an addition of O.IS lb. lead acetate per km of ore. The most economical point in grinding has been to yield a product 75 per cent of which will pass 200>niesh. The overflow from the classifiers passes now to eight 6-ft. Callow cones used as slotighing-off cones; that is, there is a turbid or slimy overflow going to Dorr thickeners and a spigot discharge forming the feed for sixteen No. 6 Wilfley concentrating-tables running 300 strokes per minute, with a f-in. throw. Close concentration is not attempted, but rather to obtain a clean concentrate product. This is put into a vacuum tank 14 ft. diameter by 3 ft. deep, and allowed to dry under vacuum for forty-eight hours. Afterward it is shoveled from the tank to a steel bin below. It is sampled and shipped by rail to the smelter. In attempting to treat these concentrates on the spot, it was found that, while a good extraction could be made with fresh solution, socm this became foul and inactive and the extraction became poor. Even where this was good, the cost of treatment would be higher than the shipping and marketing expense. Farther it was found, even when roasting, a satisfactory extraction was not attainable by cyaniding. These concentrates were about 60 per cent pyiite, 30 per cent insoluble and conteined 1.6 per cent silver and gold, or 467 oz. per ton. The tailings pulp from the Wilfley tables joins the over^ flow from the Callow cones to go to four 30 by 12-ft. Dorr thickeners (see Fig. 102). The proportion of solution to ore in this united flow, which had been added in the prior operations, is as follows : At the stamp batteries 5 to 6 parts; to correct the water at the tube-mill feed 0.8 part; at the Doit classifiers, in order to thin the pulp for proper classificatitm, 1 part; as wash-water at the Wilfley tables about 1 part. Two of these classifiers are fitted with trays, being another bottom with a central opening. Additional settling area is thus afforded, increasing their capacity by about 75 per cent. The overflow from these, the firat set, is sent to the partial precipitated solution tank 54 while the underflow of 1.26 specific gravity, is pumped to the first tank of the first series of Pachuca agitators 28, and passes through them all in series. E^h of these tanks, 15 ft. diameter by 45 ft. high, is agitated by a central air-lift (see Fig. 100). The pulp iroai the last agitator, diluted with foiir more parts of partial precipitated solution from 54 goes to a second set of Dorr thickeners, the overflow to the press-solution tank 44, while the thickened pulp is delivered to be treated in series through a second set of agitators, whereby the silver has
268 Cyanidation Op Silver Ores
been thoroughly brought into solution. The flow from the last agitator goes to 33, a filter stock-taak 28 diameter by 20 ft. high, eqiiipped with a Tcent mechanical agitator to keep the pulp from settling. The total period for agitation is forty-eight hours. By thus thinning the pulp with a nearly barren solution after it has passed through the first aeries of agitatore a large proportion of the dissolved silver ia displaced, which relieves the vacuum filters 36, from much work, and allows a change of solution for final agitation. At No. 1 agitator enoi^ cyanide has been added to bring up the solution-strength to 6 lb. cyanide per ton; also 0.18 lb. per ton of lead acetate is then put in. In the central columns of the agitators are steam coils for heating the rapidly rising slime, it having been found that by thus heating a 2 per cent better extraction is attained. The watery slime in the filter stock-tank is now ready for filtering. This is accomplished at the filter boxes or tanks, of 250 filter leaves. The pregnant Bohition from the vacuum filter boxes (now to be precipitated) ia pumped to the press solution tank 44, which is 30 ft. diameter by 10 ft. deep. It is treated in a Crowe vacuum cylinder, clarified and sent on to the partial precipitation or complete-precipitation supply tanks.
The Refinei;. — Here, for partial precipitation are three Merrill triangular presses and for complete precipitation one press. Only enou^ solution is comjAetely precipitated to a value of about 10 cents per ton, and is used for dilution at the second set of Dorr thickeners for table-wash solution and tube-mill feed-solution. The completely precipitated product carries 75 per cent silver (and gold), 7 per cent insoluble matter, 3.6 per cent lime carbonate, and nearly 5 per cent zinc and other impurities, notably selenium to the extent of j per cent. For a clean-up the product is well mixed with 2\ per cent of borax and 6 per cent each of soda and sand, and made into briquettes. These briquettes are melted in double-compartment carborundum-lined furnaces. Fig. 146, to produce a bullion containing 93 per cent silver (and gold), 2.5 per cent lead and 1.8 per cent selenium.
On about half a million tons of ore the extraction of silver was by concentration 9 per cent; by cyanidation, 84 per cent, a total of 93 per cent. Of the gold 6 per cent was recovered by concentration, 90.2 per cent by cyanidii^.
CYANIDATION OF MIXED SILVER OSES AT THE SAN FRANCISCO MILL, PACHQCA, MEX.
This plant treats ores from a number of different mines, each <rf which is different in chemical composition and requires a variation in the treatment. The first Pachuca tanks used in North America were installed at this plant.
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The San Francisco Mill 260
The average content of the ore received has been 17.4 oz. silver, and 0.08 OS. gold per ton.
The plant contains a complete sampling mill where all ores are received, crushed, automatically sampled, and diimped into bine, from which they are carried to the battery-bins in cars. The batteries consist of forty stamps, weighing 1250 lb. each, which drop 7} in. at the rate of 100 drops per minute. The ore is crushed in the batteries to pass through a No. 9 roll slotr«creen and thence falls onto eight Wilfley concentrating tables, whence the tailing runs into four Dorr claaeifiers, from which the slime and excess solution fall into four Frenier pumps, by which they are elevated to two Dorr Blime-thickeners, 24 ft. in diameter by 10 ft. deep. The sand from the Dorr clasaifiera falls into four tube-mills, 4i ft. in diameter by 13 ft. long, so placed that each mill receive the sand from one classifier. After the sand is reground in the tube-qnills, it passes into four Frenier pumps, which return it to the Dorr classifiers until the whole of the pulp has been converted into slime which passes to the two Dorr slime-thickeners previously mentioned.
From the bottom of these slime-thickeners the slime of the proper consistence for good concentration, is drawn off and fed to fifteen concentrating tables by means of which the heavy minerals which have been Uberated by regrinding, or which escaped concentration on the Wilfieys, are concentrated out of the slime. The tailings from the concentrating tables are elevated by a S-in. centrifugal pump to a third Dorr slime-thickener 24 ft. in diameter by 10 deep, whence the thickened alive, containing 1 ton of dry slime to Ij of solution, flows to the Pachuca vats for agitation, while the overflow of this, as well as the overflow of the either two slime-thickeners, flows to a tank from which, when clear, it is pumped to the vats above the mill which supply solution to the batteries. It sometimes occurs, when milling certain classes of ore, that the third Dorr slime-thickener will not have a auSicient capacity for settlement of the ore milled, so that the overflow contains unsettled slime. When this is the case the overflow from this thickener is run into a series of four masonry settling-tanks, where the slime is settled before pumping the solution to the vat€ which supply the batteries.
There are eight Pachuca tanks in this plant, each of which is charged with from 80 to 100 tons of dry slime for treatment. The slime resulting from the milling and classification, which is treated in these tanks, is of such a fineness that 80 per cent will pass through a 200-me8h screen.
The slime is elevated by pumping to a slime stor^e tank placed above the filters, from which it is fed to the filters as required. There are two filter-plants in this mill, each having a capacity of 150 tons per day. The Butters filter has seventy-eight filter-leaves, while the Moore filter has two baskets of forty leaves each. Each filter-plant is supplied with all of the
270 Cyanidation Of Silver Ores
latest iintm>veme&t8 and they give entire satisfaction. After filtmtioQ ' the Botutions are pumped to the sand filters above the precipttatitm rocKO while the slime tailing, after being filtered and washed, is discharged into the tailing-dam. The solution from the taihng-dam, after settlement at the elime, is pumped to a tank above the filters for use in them as wa^ water.
The precipitation room contains ten zinc-boxes made of sheet sted, each 15 ft. long, 3 ft. wide, and 2^ ft. deep which are divided into fire compartments each. The precipitation of the clarified solutionfi fnnn the sand filters is almost perfect in these boxes, as the sohitioos entering the boxes assay fnun 100 to 250 gm. of silver and from 1 to 3 gm. of gold per ton, while the precipitated solutions leaving the boxes do not ctary more than 1 gm. of silver and 0.1 gm of gold per ton.
The cakes of precipitate from the filter-pross are dried until they coo-tain from 5 to 10 per cent of moisture. Then, after brefiking up the larger tumps, a mixture is made, conaiBting of precipitate 81 per cent, sodium carbonate 7 per cent, borax glass 10 per cent, quartz tailing, 2 per cent.
A graphite crucible. No. 300, is placed in the coke furnace, and, as soon as it begins to heat up, is filled with the mixture of precipitate and flux, and melted in the usual way. Total treatment costs are SI .93 per ton, with recovery of 92.4 per cent of the silver and 94.5 per cent of the gold.
THE WAim ORAin} JUNCTION HILL, WAIHI, NEW ZSALAHD
The ore from the mine coneiste essentially of a gangue of quarts and calcite, with 8 to 10 per cent sulphide — pyrite, sphalerite, galena, cbalcopyrite, and traces of arsenic and antimony. In sloping, this is mined with the well-defined lode, thus bringing into the mill at times a considerable quantity of hard country rock. The gold exists in an exceedingly fine state of divisicHi, consequently rendering it necessary to grind very fine to obtain a satisfactory extraction. The ore averages about $8 gold and $1 sitver per ton. There are forty 1100-lb. stamps fed by a hanging type of improved Challenge ore-feeder. Cyanide solution of 0.10 per cent is added here in the proportion of 10 of solution to one of ore, also lead acetate equal to 0.5 lb. per ton of ore.
From the stamps with a duty of 7.6 tons a day the pulp flows to three elevator wheels,' by means of which it is raised to a series of conical boxes or spitzkasten. The overflow from these boxes passes on to the WUfley tables. The underflow, after passing through the tube-mills, is again returned by the same elevators to the conical boxes. The Wilfley tables are used merely as classifiers. The discharge is taken from 18 in. along the
' The t&ilings wheel, though of greater hrst cost, is reliahle and the cost of rapaits b
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THE WAIHl GRAND JUNCTION MILL 271
rede, and the tables are so set that their overflow is practically fine enough for ecoiuHnic treatment. The heada, consifrting of the lai^^ particles of
Fia. 148.— Flow-eteet of Waihi Grand Junction Co.'s Mill.
concentrate and sand, are sluiced to belt elevators, which raise and die- ^hargR into the feed-cones of the tube-mills.
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272 Cyanidation Of Silver Ores
There are fourteen tube-mills, run at 27 R.P.M. The average Kfe rf the Imers is e^teen months. The angle shoes are alao made of. hard cast iron, and have an average life of ten months. The pulp is fed into the mills by means of injectors. The feed noazles are Ij in. made diametn and discharge into larger cast-iron pipes which are bolted on to the end tA the mill. The clearance between these pipes is | in. The ratio of sohitiMi to ore in feed is as 1 ^ to 1. Cost of tube-milling is about 28 centa per t<m.
The overflow of the Wilfley tables flows to a spitzlutte, the spigot product of which is returned to the tube-mills by a centrifugal pump. The overflow, consisting of solution and slime in the proportion of 10 to 1, flows to the settleiB or thickeners, lime being added to the launder that conveys it. Of the overflow of clear solution, from the settlers, 75 per cent flows to the strong solution sump and 25 per cent to the strong solution clarifying tanks. The pulp is drawn continuously from the bottom of the settlers, and is of a consistence of 1) of solution to 1 of slime. This is pumped into one of a series of 12 flat agitators, each 22} ft. by 6 ft., provided with four arms revolvii^ at the rate of 4 R.P.M. These agitators are used as storage tanks. When three are full they are dkcharged by a pump into a tall tank, 55 ft. by 13) ft. Agitation is by compressed air at a pressure ot 38 lb. per square inch. Each tank has a capacity of 250 tons of pulp. Agitation is continued in these tanks for an average of 18 hours and gives a further extraction of 12 CRnts per ton. The strength of cyanide is maintained by a continuous flow of strong stock solution, which is added to that supplying the mortar boxes. This is the only place where cyanide is added. The consumption of KCN is 1.258 lb. per ton.
From the tall tanks the pulp gravitates to the Moore filter plant. Each basket is composed of ten framra, 16 ft. by 4 ft., giving a total filtering area of 1280 sq. ft. per basket. The weight of slime cakes in each basket is equivalent to five tons dry weight; seventy-five minutes is required for formation with a vacuum of 25 in. of mercury. The loading tanks are fitted with conical bottoms, and an air-lift is anwiged to prevent settlement of slime. The basket is raised by an overhead traveling crane, electrically driven, and transferred first to a wash-tank <J weak solution (0.04 per cent KCN), where the charge is washed for forty minutes, then to a water tank, where a further wash of water is drawn through for ten minutes. The basket is now raised and suspended over the discharge tank, drained thoroughly, and the tailing sampled for assay. When the vacuimi is destroyed the charge falls on to the revolving arms in the tank, and is sluiced away with wa1«r jets to the sludge channel. '
The cost of treatment, which includes agitation and vacuiun filtration, is 64 cents per ton.
The vats, receiving the solution from the slime treatment, are provided with filters to prevent slime from passing into the zinc-boxes. From these
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THE WAim GRAND JUNCTION MILL 273
vats the gcdd-bearing solutions 6ow tbro\^h meters, which register the tonnage, to the launder feeding the extractor boxes. Theie aie tw»ity-two boxes, seventeen of which are used for strong solution (0.10 per cent KCN) and five for weak stdution (0.0 4 per cent KCN). The strong solution boxes h&ve eight compartments, with total available zinc space of 740 cu. ft.; the weak solution boxes-have a total zinc space of 145 cu. ft. All are provided with a screen of 4-mesh near the bottom of each compartment. The average rate of flow of solution through the strong boxes ia 1.4 tons per cubic foot of rinc per twenty-four hours, and, through the weak, 1.1 tons. Precipitation of the metals from the weak solution is assisted by first dipping the zinc in a solution of lead acetate. At the head and tail of each set of boxes is an automatic solution sampler. These samples are assayed daily, and from the results, together with the tonnage passed through the boxes as registered by the meters, the amount of buUion precipitated can be calculated. The bulk of the precipitation takes place in the first two compartments. The first four are cleaned out every seven days, and all the compartments every twenty-eight days — that is, at the end of each period. When the boxes are cleaned the zinc is washed in the first cells, the precipitate is drawn off by means of a suction hose into a receiving cylinder, thence into vacuum filters. The short zinc, which passes 4:-me8h and remains on 2S-mesh, is treatedwith 1 to 6 per cent sulphuric acid in vats fitted with revolving arms, which agitate the charge, which, after being well washed with hot water, is dried with the balance of the precipitate. The precipitate is dried in cast-iron ovens without stirring. The slime and fluxes are mixed in a small tube-mill, which is entirely enclosed and dustproof, and when thoroughly mixed it is discharged into a truck. Melting of the precipitate is conducted in No. 120 graphite crucibles. Kerosene fuel is used, and, with four furnaces, two men melt and refine 1100 lb. of precipitate in eight hours. The buthou and slag are poured together into conical molds, and, when solid, the bullion is detached and re-melted in the same crucibles and poured into bars weighing about 1100 oz. each. The slag is crushed by stampers through 25-mesh screen, passed over blanketing to retain prills of buUion, and collected in settling boxes. It is afterward aiivdried, bagged, and shipped to smelters for treatment. The cost of preciintation and melting is 5 cents per ton of ore crushed in mill.
Two Balback tilting furnaces, using coke fuel, have been installed, and are used for melting the precipitate, the kerosene furnaces now being used for remelting the bullion into bars.
The total cost of milling and treatment is $1.50 per ton.
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Cyanidation Of Silver Ores
ICnXINO PRACnCB AT COBALT, ONTARIO
There are two grades of ore, the high and low grades, each treated by a different procesB. The treatment of the hi^-grade ore i& givim on page 243; the milling of the low grade is described below.
The On. — This is ocmplex, containing chiefiy native kIvbt occurring in particlea entangled in a mixture of nickel and arsenic minerals (smallite and uiccotite) and calcite. It is classified in two grades, the " hi^ grade," chiefly of native silver of 2500 ok. per ton, and the " low-grade," having native silver, also considerable sulphides and antimonidea, whose silver is difficult to recover.
Prelimioaiy Treatment — At the washing plant, as shown on the flow sheet, Fig. 149, tiie mine ore is dehvered to a trommel 40 in. diameter by 10 ft. kmg, where with the addition of water it is washed and separated into two products, the overmze going to two seta of trommels to yield producta for two sets of jigs. The jigs yield a high-grade head whidi joins the corresponding product from the 30-in. picking belt to go to the high-grade ore mill. This ore amount^ to about 9 per cent of the total, and its treatment in the high-grade mill, as indicated at the ri^t side of the flow-fiheet, is described under head of " amalgamation of silver ores," page 243. This preliminary treatment lessens the period of agitation necessary with cyanide solution.
Many experiments involving the preliminary reduction of telluride gold ores in caustic soda solution with aluminum and zinc, as well as cathodic reduction in both caustic soda and salt solution, have shown that with cheap fuel and efficient roasting this form of preliminary treatment would present no advantage. The gold telluride compounds are in general more difficult to reduce than the sulphide and sulpho-antimonide silver minerals.
To go back: the imdersize from the last trcHumel, together with ih& jig tulings, is elevated to a dewatering trommel, 30 in. diameter by 6 ft. long. The oversize joins the ore from the picking belt destined for the low-^rade mill, while the underaize after dewatering is Beat to join the battery pulp in the same mill.
THE HIPISSmG CO.'S " LOW-ORADB " HILL, COBALT, OHTAKIO
The ore, of about 26 oz. silver per ton is treated by all sliming cyamding. The left-hand portion of Fig. 149 gives the flow-sheet of this miU. There are forty heavy stamps of 1400 lb., each battery o{ ten stamps being driven by a 40 H.P. motor, wetrcnishing throiigh a 2-to 3-meah screen and using for each ton crushed 7 tons of a solution containing 0.7 lb. of caustic soda per ton. This caustic soda addition is necessary as a preliminary to the desulphurizing treatment to follow.
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The Nipissing Low-Gbadb Mill
tocCniKitvi Fla. 149. — Flow-sheet of Nipisaing Co.'b Mill.
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276 Cyanidation Of Silver Ores
ClEBsificatioii and Fine Grinding. — For high extraction the ore must all be gnnmd to pass 200-mesh. The bafteiy pulp is classfied in a first Bet of four Dorr claesifieis, the slime overflowing direct to one or other of the slime-collecting vats in the cyanide plant, the oversize goii^ to two coaise-grinding tube-mills. The discbarge from these mills flows into a second set of Dorr classifiers whose overflow joins that of the first set at the 8lime-«ollecting vats, and the sands are elevated to two fine-grinding tube mills. The discbai^ from these flow into & drag classifier, then back to the second set of Dorr classifiers all in closed circuit with the fine^rindiog
^ Cyanide Plant. — This is in the largest building, where the vats are placed in two rows, and, with the exception of the solution vat, fitted with mechanical stirrers driven by a 125 H.P. motor at 8 R.P.M. The working load in each vat is 140 tons of dry stime with 280 tons of solution. The slime-collecting vats are dewaterers delivering a clear overflow to a lower crushing solution vat and thence by pipe line marked " caustic soda sohition," to the upper crushing vat for re-use at the battery. When one collecting vat is full of a charge the pulp flow is switched to the other. The charge is then agitated for an hour mid the thickened pulp, conasting of 1.5 parts of caustic soda solution to one of the pulp, is pumped to desulphiuiEing treatment.
The Wet Desulphurizing Process. — ^This breaks up the refractory silver minerals, the slimed pulp being brought in contact with aluminum in the caustic soda solution in a tube-mill and the silver left in the spongy metallic state unenable to cyanide treatment. To do this the collecting vat imlp passes to a desulphurizing tube mill, revolving 10 R.P.M. where it comes in contact with a load of about 4000 tb. of alimiinum pieces of 1)- to 2-in. cubes. The pulp discharge from the mill gravitates to the desulphurizing vat adjoining. In this tank 34 ft. diameter by 13 ft. deep and lined with ahmiinum in plates, the pulp receives mechanical agitation for a period of twenty-four to thirty-six hours. In order to keep the mill cyanide solution in balance it is necessary to eliminate as much as possible of the caustic soda solution from the desulphurized pulp. This is done in a 60-leaf Butters vacuum filter, marked " alkaline filter " in the plan view. This preliminary treatment lessens the period of agitation necessary with cytuiide solutions.
The filtrate passes to the crushing solution vat while the thick slime is pumped to one of the seven cyanide treatment vats fitted for mechanical agitation. There it is treated in charges of 130 tons of dry slime to 260 tons of cyanide solution of 0.25 per cent (5 lb. per ton) with 0.2 per cent alkali. An air lift is operated constantly during agitation, the pulp being drawn off from the bottom and discharged into the top of the vat and treated at the same time. After agitation the chai^ is allowed to settle
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The Nipissing Low-Grade Mill
so that the clear solution may be decanted to the " pregnant solution vat." The pulp is then stirred and pumped to a 34-ft. cyanide stock-pulp vat, where it must be kept agitated until drawn off to the 80-leaf Butters vacuum *' cyanide filter." The clear solution from the filter is delivered to the " pregnant solution " tank, the residue slime is discharged to the residue dump.
The pregnant solution is now ready to go to the precipitation department at the righb of the tube-mill house to be subjected to aluminum precipitation treatment for twenty-four hours.
The cost of treatment in 1912 was given as $3 per ton. The cost of the plant ^M.OOO to treat 244 tons daily.
CTANIDATION OP SltVSR-BBAIUHG CONCBNTRATBS
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i
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17
C. Tube-milled in 1 lb. KCN s(Jution; aptated m 6 lb. KCN for ten to fourteen days, with two decantations; filter pressed and precipitated by zinc dust; 2 lb. lead acetate and 1 lb. mercuric chloride per ton: of concentrates. :
D. Tube-milled and agitated in 3 lb. KCN solution far eight to ten days.
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Chapter Xxiv
PARTmG SILVER-GOLD BOIIJOH
Parting Silver-Gold Ingots Or Bars With Acids
Tbe bars from reduction works commooly contam gold and diver alloyed with copper, but sometimeB also sine and lead. It is customary to remelt the bars and assay them, buyiug them on tbe result of the assay.
The bars are parted in nitric or sulphuric acid. Sulphuric acid, being cheaper, is the acid commonly employed.
Bars containing a lai^ proportion of gtdd are inquartated by melting them with silver in order to decrease the ratio erf gold to silver to at least 1 to 2.5, otherwise the acid fails to attack the silver. In parting with sulphuric acid the copper should be less than 10 per cent, but in nitric-acid parting more than 10 per cent is allowed. To adjust this percentage, bars low in copper are melted with those high in that metal.
Nitric-acid Parting. — ^This method, still practiced at the United States Mint, Philadelphia, is an efficient way of parting, especially on a small scale. The bars having been melted and proportioned as above described, the molten metal is granulated by pouring it into a tank of water. The granulated metal is transferred to porcelain, ^ass, or platinum vessels, and treated with nitric acid, 1.20 sp. gr., until action ceases. The solution is allowed to settle, then is decanted, and fresh acid is added for tbe purpose of dissolving the remaining traces of silver. This is again decanted, and the gold residue is washed thoroughly with hot water. It is then ladled out, drained, dried, aad mixed with a Uttle flux, and melted in a graphite crucible.
From the decanted silver solution the metal is precipitated by adding common salt. The silver chloride thus formed is washed, thoroughly granulated, zinc is added to reduce the silver to metal, and the zinc chloride resulting from the reaction is washed from the precipitated silver. The silver is pressed into cakes, melted, and cast into ingots for bar-silver.
Sulphuric-«cid Parting. — Since for this method of parting there should be less than 10 per cent copper present in the gold-silver alloy, and not less than 2.5 pari» silver to 1 of gold, the bars to be parted that exceed Uie reqxiired limit, are bo selected and melted with others as te afford tbe requu^d proportion. The melted metal is cast into flat ingots and parted in this form.
27S
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PAHTINQ SILVERrOOLD BULLION 279
The ingots are pl&ccxl in a cast-iron kettle, covered with a sheet-iron hood that is connected with a chimney bo that the acid fumes from the kettle are carried away. Here they are ttcated with sulphuric acid of full strength (66° B4.)- When acti<ai has ceased the solution is allowed to settle, after which the clear supernatant part is decanted, being drawn off by a lead-pipe apron mto a lead-lined pTecipitating tank. The residue in the kettle is treated six or seven times with fresh boiling acid. In this way the silver completely dissolves, the acid solution being removed after each treatment. The brown gold remdiie is finally boiled with water, being heated and agitated by hve steam from a ppe inserted in the water. In this way the gold is " sweetened." The residue is removed from the kettle, dried, melted in crucible with a httle borax for flux, and cast into a bar of gold 999 ane.
- The acid solution from the kettles, which flows to the precipitating tank, is diluted with water, and the ulver is precipitated t^ hailing copper plates about 1 in. in thickness in the solution. The copper replaces the silver in the acid solution, which becomes blue in color. When precipitation is complete the clear solution is decanted, and the cement silver at the bottom of the tank is washed with hot water to remove the acid coppersohition. The " cement " silver, or precipitated silver, is removed to a box, then pressed into cakes or cheeses in a hydraulic press. Thus compressed, it is ready for melting in plumbago crucibles after adding a httle borax flux. In large establishments the silver is melted in a small reverberatory furnace where it can be conveniently fluxed, skimmed, and ladled into bars for the market. The bars weigh 35 lb. or 500 oz. each. On refining, each bar is marked with a number and the exact weight and fineness, and the name of the refinery that produced it.
BLECTROLTTIC PARTIHO OF GOLD FROM SILVER The U. S. Mint, San Francisco, receives bulUon of 200 fine or over of gold and silver. Gold bars of 900 fine or over are treated by a gold-refining process, while others are melted tc^tber to be cast into cathodes 600 fine in ulver, 300 in gold, and the retnaining 100 in base met&ls.
Silver Refining. — The anodes, 3} in. wide by 8j in. high by ) m. thick, are hung in eulibenware cells t(^ether with thin sheets of silver for cathodes, using for an electrolyte a nitrate of olver solution that carries 3 per cent silver and 1\ to 2J per cent of free nitric acid and using a current of 0.8 ampere per square foot. The pure.silveT collects on the cathodes as a sponge deposit. Part falls to the bottom of the cell as slime. From time to time the anodes are taken out and the sponge jarred off. This and the slime are melted to produce a gold bar. The pure silver collects on the cathodes in crystalline condition and they are lifted out daily to remove the silver. The final product is melted into silver bars.
280 Parting Silver-Gold Bullion
Gold Refining. — The anodes of the same size a^ the Bilrer odbs should be at least 900 fine in gold and leas than 70 fine in silver. Hiey are hung in porcelain cells with thin pure gold sheeta for cathodes and in an electrolyte consisting of gold trichloride having 7 per cent gold and 10 to 12 per cent <rf free hydrochlorine acid. A high current (90 amperes per square foot) is used. The cathodes when built up to 160 oz. each are rranoved and cast into ingots of 999 fine.
The spent electrolyte is replaced by fresh. When it gets impure it is separately treated to recover some silver and the base metala.
PRICES Aim COSTS
The prices of silver ores sold to custom smelteries are given under head of Uie schedule of silver-lead ores being for a dry ore.
Silver. — At New York the quotations are on silver bars, per troy ounce of silver, 1000 fine. It takes 14.58 troy ounces to make 1 lb. avoirdupois. London prices are for sterling silver, 925 fine. The value of the pound steriing is also given, so that with the London quotation, we may compute tiie equivalent price in cents there. Let us aay that sterling exchange is J4.88, and that silver is selling at 25d. per sterling ounce, we have then:
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Part Iv Iron And Steel
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Chapter Xxv
Iron Ores And Their Smelting
CLASSIFICATION AlfD OCCnitIt£NCB OP QtOH ORES
Tbeae, Uie oxides and carbonates of iron occur accompanied by earthy mineralB forming tbe " gangue." Only those are regarded as iron ores that contain sufficient iron to make the recovery of the metal profitable. In this discuseicai of iron ore it is to be understood that reference is made to them as smelted for the uxin they contain, and not as a flux as used in silver-lead and cc^per smelting where the iron enters a slag to be thrown away.
There is a general division of iron ores into two classes, viz., Bessemer and non-Bessemer. This arises from the fact that a steel containing more than 0.10 per cent phosphorus is brittle, and hence, in the acid-Bessemer steel-making process, tbe pig iron used must have rather less than that quantity. Now since about 2 tons of iron ore make a ton of pig iron, the iron ore should, for safety, not contain more than 0.045 per cent phosphorus. At about that limit the ore is called non-Bessemer. In the duplex process of steel making this distinction has become of little importance, since a steel from the Bessemer converter, high in phosphorus, can be freed from it in the open-hearth treatment that follows.
Iron ores, especially in the United States, are oxides. They may be divided into hematites, magnetites, and brown iron cnes. The carbonates are there found sparingly.
The Hematites.— The iron in them exists as FeaOa (70 per cent Fe). Of iron ores red hematite is the most desirable. Most of the Lake Superior deposits are of this variety. The ores from the Michigan and Wisconain districts or ranges are called " old-range ores," Much hard or lumpy ore comes from these ranges while from 1^ Minnesota ranges much of the ore is soft.
Following we give the c(»npoation of seme of these ores as shipped to the iron furnace in their natural (or moisture-containing) condition.
But Mesabi ores will vary from 39 to 67 per cent m iron. The poor csta are now ccmoentjated to bring them up to shipping grade.
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Iron Ores And Their Smelting
H,o.
F.
So,
a.
P.
Mn-
2.fl 7,5 11,0 3.5
Gogebic, Wia
Meaabi, Minn.:
Non-BMsemer...
An average analysiB of Lake Superior shipping ores in 1919 ^ve H2O, 11.3 per ceat; Fe, 58.4 per cent; SiOa, 7.7 per cent; S, 0.06 per cent; P, O.Od per cent; Mn, 0.7 per cent, the bulk of the ore being (rf Bessemer grade.
Among the ores of the United States we should also mention the Alabama beda, and those of Colorado and Wyoming in the West.
There is a red hematite at Sunrise, Wyo., which carries Fe, 62 per cent and beds at Orient, Colo., of easily reducible limonite of 50 per cent.
The Alabama ores may be divided into three varieties : First, the brown ore or limonite, averaging Fe, 51 per cent; P, 0,4 per cent and S, 0.10 per cent; second the soft hematite of Fe, 47 per cent and with as much as 17 per cent in silica; and third the hard red carbonate ore of Fe, 37 per cent; SiOa, 13 percent; P, 0.37 per cent CO2 and 12.2 percent. Theyarehard, heavy and hence, usually nearly black in color.
Magnetite occurs in large deposits in Sweden, and in various parts of the United States. While some of the beds are rich, many contain no more than 40 per cent iron and carry so much silica that the fiuxing and smelting is not profitabli!. Much work has been done 'n the concentration of these ores, both in Sweden and the United States. In New Jersey extensive beds occur that have been utilized by Edison for the production of a high-grade ore on a commercial scale. He has mined the deposit, crushed and concentrated it, and made it into briquettes that contain as little as 3.3 per cent alica and 0.04 per cent phosphorus, and as high as 67 per cent iwai. The enterprise, however, could not continue at a profit in competition with fore^n ores from Cuba and Spain. Cuban ores include magnetites of Fe, 50 per cent; SiO, 10 per cent; S, 0.37 per cent; Cu, 0.10 per cent. The Spanish spathic ore has HaO 0.91 per cent; Fe, 48 per cent and SiOs, 10 per cent. Some of the New York beds near Lake Champlain have been considered valueless on account of the presence of titanium, it having been asserted that this element produces an infusible sticky slag. This, however, has been proved to be unfounded, and it should not prevent their use as a source of iron. It is to be noted that the famous Iron Mountain, Missouri, is a deposit containing 31 per cent iron and 6 per cent titanium oxide, but the deposit
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Iron Or£S 285'
is not now worked. In Pennsylvania the Cornwall beds are the most important, and yield a pig-iron carrying not more than 0.04 per cent phosphorus. The ore runs 2.5 per cent sulphur, and about half of this is removed by kiln-roasting before smelting. It contains also copper, which will be found in the pigto the extent of 0.5 to 0.75 per cent. "Hiia does not matter in the finished product, but, if the pig-iron is made into steel, the copper causes " hot^hortness " or brittleness when hot, thus causing imperfections when rolled into shapes. The average of ore mined is 40 to 42 per cent iron and 20 per cent silica.
At Fierro, N. M., is a large deposit of bard magnetite riuining up to - 61 per cent in iron.
The Brown Ores. — -These are hydrous sesquioxides and may, when pure, be represented by the formula Fe20a+3H20, equivalent to 60 to 68 per cent iron. Limonite, locally known as bog ore, a brown hematite, when roasted has this water of combination expelled, changing then to true hematite. Oolite is a variety that exists in the form of grains or nodules, and contains silica and Ume. When sllicious, as in places in Alabama, the ore is well-nigh worthless, but when limy, as in the Minette region of Alsace and Lorraine, the ore is setf-fluxing, that is, the lime will Sux the silica that the ore contains. A typical Minette ore carries Fe, 38 per cent; SiOa, 9.2 per cent; CaO, 12.1 per cent. In the United States these ores occur parallel to the Appalachian range from Pennsylvania into Alabama, and on both sides of the Mississippi, in Tennessee and Missouri. Due to the presence of much phosphorus these ores are of non-Bessemer grade. They will vary from 40 to 50 per cent in iron, 5 to 20 per cent in silica, 0.05 to 0.4 per cent in phosphorus, and from 0.3 to 2.0 per cent in nu^- nesia.
Carbonate ore, dderite (FeCOs), as a pure mineral contams 48.3 per cent iron. The varieties are spathic, black band, clay-band, or clay ironstone. It is often roasted to expel the moisture and carbon dioxide before going to the blast-furnaces. In England it forms the well-known clay iron-stone of the Cleveland district, but in the United States, though widely distributed, it is too low in grade to be used in competition with ihe abundant rich ores.
At Eisenerz, Styria, is a celebrated deposit of siderite. The ore, which averages 39 per cent iron, is worked in vast open cu1«. The Sparudi spathic ore has HaO, 9.1 per cent; Fe, 48 per cent and Si02, 10 per cent.
ROASTING mow ORES
In order to expel moisture, carbon dioxide, and sulphur, and to render the ore more porous, and so more susceptible to reduction in the blast-furnace, aa well as to decrease its weight, iron ores, especially carbonates, are
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286 Iron Ores And Their Smelting
often rotated, preferably In kilns of which F^. 150 is an example. The ore is charged to the kiln in layers alternated with sufficient fine coal, so that t^e roasting heat is maintained and the kiln is kept quite full. Peep-holes or openings at the boshed part of the kiln are for obeervatioa and for loosening the charge by aid of a bar as needed. Ore is unloaded, as shown, from cars into the kiln. The roasted material runs out at the bott^nn as
Vertical Section
Fto. 150. — Gjiera Calcining Kiln.
fast as it is removed. Air is admitted to the midst of the charge under the cast-iron hood, at the apex of the central cone.
THE AOOLOHERATIOn OF FINE ORES
In modem blast-furnaces, due to their high stacks and heavy blast pressure, it is difficult to smelt a high percentage of dusty or finely granulated ore, such as those of the Mesabi range. Where 60 or 70 per cent of such ores must \m used there i.'^ a heav>' flue-dust loss, scaffolding of the furnace, and frequent explosions. The fine ore, descending more quickly than the rest of the charge, is imperfectly reduced, causing disturbances in furnace operation and the production of " off-iron," that is of pig iron of other than the expected grade. Fine ores, therefore, ought to be put in lump form. This may be accompli^ed :
(1) By noduUzing in revolving kilns or cylinders similar to Fig. 72 but longer. For such ores as shrink much in roasjing the method is valuable, but the product is seldom uniform. . (2) By blast-roasting. This is done on a Dwight-Lloyd sinter machine,
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PIQ-mON SMELTINa 287
as described on page 112. The product ia poioua and well mntered to blast-furnace smeltisg.
(3) By briquetting. This is dtoie vith or without Mnding material, and is usually followed by the heating of the briquettes to resist breaking and the heat in the blast-fumace. Care must be taken in making them that they retain their porous character.
SKHLTinO FOR PIO-IROH
Outline of fbe Process. — The operation is conducted in a furnace, often 100 ft. high, filled with a mixture of coke, iron, ore, and limestone. Superheated air is blown in at the bottom. The coke is bumed to maintain a high temperature in the furnace and to reduce the iron m the ore to the metallic form as pig iron. The pig inm collects at the hearth or bottcon
rVia. 161. — Three Traversing-bridge Tramway with 6)-ton Gnb-bucket; Storage capacity, 500,000 tons.
' of the furnace, and is removed from time to time. The gangue, or'ailiciouff part of the ore, is fiuxed with limestone, and produces a worthless slag, or cinder, which is also removed (tapped) as it accumulates in the furnace.
IKON BLAST-FUItNACB Ain>. PLANT
Blast-funiace Plant. — ^Fig. 152 is a view of an iron blast-furnace plant for the manufacture of pig iron from iron ores. In the foreground is a cylindrical furnace-stack 100 ft. high, immediately in front of which is the forked " down-comer " (see 39, Fig. 155), a large pipe that conveys the smoke from the stack near the top downward to the fiue-system that carries it away. In front of the down-comer is seen the inclined hoist for the " stock " or the materials that are put into the furnace. At the middle of the illustration are the four cylindrical " stovea," as high as the furnace, used for prebcatii^ the au-blown into the furnace, while the highest stack
288 Ihon Ores And Their Smelting
behind them draws away the gas from the stoves. In front of the four stoves b the blast-main, a pipe 5 ft. diameter by which the air is conducted to the furnace. At the left of the stoves is the building (not shown),
Fia. 152. — Blaat-fumace Plant'.
that contains the vertical blowing engines by which air, under 15-lb. prea-' sure, is deUvered through the stoves to the blast-furnaces. Fig. 160 represents a vertical blowing engine.
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The Iron Blast-Furnace Plant 289
The general arrangement about the furnace is understood from the elevation Figs. 151 and 153 and the plan Fig. 159. The blast-fumace 100 ft. high is at the right. It is served by an inclined hoist, one skip of which is in position in the pit ready for loading, the other in discharging position at the top. Within the stock-house, at the left, are three standard-gauge tracks on three levels. The upper one at the left is for the hopper-cars that deliver iron ore and fiuxes to sloping-bottom bins beneath, shown in section.
The second one leads to similar bins (not shown in section), and the third to the floor on th^ ground level. On this level is a charge-car, elec-
Kio. 153. — Section of Blast. Furnace, Showing Fillinn Arrangement, Bins and Ore-bridge.
trically driven, with a weighing attachment that can be brought to any bin to receive a weighed amount of stock. Thp load is then transferred to and discharged into the skip. In case of accident to the charge-car, or any trouble at bins, the furnace can be supplied by the use of hand-barrowS or buggies, taking the stock from the piles that have been made beneath the third track. Hoisting is done by a hoisting engine set well out of the way at the top of the stock-house. Just beyond and at the right of the furnace is the cast-house, where the molten iron is molded into pigs when a •cast is made.
Iron Blast-furnace Plant at a Lake Port. — We give in Pig. 151 a transverse elevation of a furnace plant as arranged at a lake port where the ore has been stored for use in the winter months and where a
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290 Iron Ores And Their Smelting
traveling craae is used to reclaim it. The aame crane has also been used to unload from the ore boat at its other end. The ore is taken frcon the pile by means of a lO-bon grab bucket which loads it Into cars standing upon tracks, these set so they can readily discharge into the feed pockets
Fifl. 154.— BUst-fumace with Automatic Chargini;.
of the blastfurnace. Coke is brought in directly from the coke ovens by car. The feed pockets, as shown, are semicircular and both coke and ore are withdrawn from them much as shown in Fig. 154. It will be seen that the charge-car, which is self-weighing, runs imder any c^ the feed
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The Iron Blast-Furnace 291
pockets, and the materials weighed into it are then taken to the feed skip of the furnace.
The Iron Blast-furnace. — Fig. 155 is a furnace, shown in part section, and part elevation. It is circular in croeB-eection.
Beginning at the bottom there is a heavy foundation of concrete and firebrick upon which rests the hearth 15 and columns 4 which support the
Fio. 155. — Blaat-funmce, Detailed Section.
Upper brickwork that constitutes the shaft of the furnace. The hearth or crucible (14^ ft, diameter by 9J ft. deep), that contains the molten iron and eiag, extends from the foundation to a height slightly above the tuyeres 22. The bottom and walls (see Fig. 156) are of firebrick, high in alumina like V of page 34, but soft and porous. Near the bottom is the iron-tap 27, through which the molten pig iron is withdrawn when a quantity has
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292 Iron Ores And Their Smelting
accumulated. At 23 is the cinder-notch or tap by which the slag or cinder is drawn oB. The crucible is murounded by a hearth-jacket of ateel plates,
Pig. 156. — Blast-fumnce Hearth and Bosh.
cooled on the outside by sprays of water that play against it, cooling and protecting it and the brickwork lining from the corrosive action of the mol-
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The Iron Blast-Furnace 293
ten alag inside. Air, imder a pressure of 5 to 14 lb. to the aqiiare inch, enter through the tuyeres 21, which have projecting nozzles 22, as more fully shown in Fig. 156. C&ie is taken to withdraw the slag before it reaches the level of the tuyeres, for it would enter the openings and cloee them. Of these tuyeres there are six. The air is supplied through the tuyerestocks 33 from the bustle-^pipe 13, which encircles the furnace, and connects with the blast-main supplying air at the temperature at a red heat from the stoves. The bustle-pipe, tuyere, and tuyere-stock are riiown in the section Fig. 156. The bosh, or that part of the furnace that widens from 14 ft. 6 in. at the hearth to 22 ft. in a vertical distance of 13 ft., is also shown. It is in the region of the bosh that the formation of the slag occura, and the brickwork of the boeh is subject to a slag^ng and scouring action that tends to attack and destroy it. To prevent this, hollow water^ cooled bosh-cooler plates are laid in the brickwork of the bosh, making rings around the furnace at n^ly every two feet vertically. The slag cuts into the brickwork nearly as far as the inner ends of these plates, but the circulation at water within them protects the adjacent brickwork from deeper corrosive action. The bricks of the bosh are a little harder than in the crucible, but are high in alumina and porous.
That both the bustle pipe and the tuyere are lined with firebrick is well indicated.
The shaft, or main brickwork structure of the furnace, is carried by the cast-iron mantel 5, resting upon the cohunns 4. It extends from the top of the bosh to the throat at 9. The upper part ot the furnace ie closed by a bell 47 (Fig. 154 shows a double bell), and the gas escapes at the side through the down-comer 39. The in-wall 69, is of a hard firebrick U\se IV, page 34,while the main portion is of common brick and is sheathed with a shell 46, of steel plates.
The tendency in modem practice is to insert cooler-phttes extending through the inwalls from the mantle upward, carrying these higher and higher as the necessity for protection demands.
When in operation the furnace is kept full to a level just below the outlet to the down-comer. This level is known as the stock-line, and the furnace at this point is 15 ft. diameter. As the stock smelts and sinks, charges are introduced and the stock-line is maintained at this level.
Ordinarily the top of the furnace is kept closed by the conical bell 47, which is suspended from the ends of the counterweighted beams 55. The beli closes the bottom of a circular hopper 48, into which the charge in this particular furnace is supplied by bu^es brought up by the elevator to the upper or charge-floor of the furnace or " tunnel-head," as it is called. To drop a charge into the furnace, the outer end of the lever is raised by the piston-rod and piston of the air-cytinder 60. The belt thus lowered permits the charge to shde into the furnace, after which it is immediately raised
284 Iron Ores And Their Smelting
to close the opening and stop the outward rush of amoke and gas that mainly escape through the hood 61. The gas, containing dust from tiie charge, passes off by the brick-Uned down-comer 39 to the dlist-c&cher 40 (where a part of the dust eettles), and by the goose-neck pipe 41 to an \indergTound flue that leads to the stoves luid boilers where the gas is burned. Rising fnHn the down-comer is the bleeder 37, that ie used when it is desired to relieve the top pressure of the gas rising from the chan^ It is occasionally used. Other openii^p are provided closed by weighted doors, called explosion doors, so that in case of a slip or fall of material due to the giving way of a scaffold or handing up in the furnaces, relief is given to the high pressure of gases suddenly released. At many furnaces the stock is raised in hand-barrows or charge-buggies to the fumace-top <xt tunnel-head 51 by means of a platform hoist.
In Pig. 154 is shown the present method of charging with the inclined hoist. A double bell is used to prevent the escape of the gas. The charge is dropped from the hoist into the upper hopper, where it is retained until the lower hopper is empty. The smaller upper bell is then lowered and the charge slides from the upper into the lower hopper, while the upper bell is closed. The hopper is then ready to take another charge. The charge in the lower hopper, when needed, is dropped into the furnace by lowering the lower bell. It shoes outwardly to the walls, forming a ring or ridge, the stock in the middle being a httle lower than at the sides.
However, the skip, dumping only in one direction, is apt to deliver the coarser ore to the opposite aide of the furnace, so that the Uast comes up more freely there, producing greater heats and a " hot spot " on that ade. To overcome this trouble dietributors are used where the charge in the bell is rotated through a varying arc which dehvers it successively to the different segments of the fmnace. Just beyond and below the lower bell is noticed the oval outlet to the down-comer. The stock-line must be kept below this.
The skips of the hoist run in balance and are charged as follows: The chai^e-car on the ground level is run to the chute of an iron-ore bin to receive the required we^ht of ore. It is moved to the limestone \an beyond to get the needed quantity of limestone, and then to the skip standing below in the charge-pit, where it is discharged. The skip is next hoisted and dumped, while the empty one is in position to take the load of coke. After elevating the fuel, a charge of ore and Sux goes next. These chaises alternate in the furnace and form layer upon layer.
The dimensions of a blast-furnace are limited. The considerations are as follows: The hearth should be not more than 15 ft. diameter lest the blast fall properly to penetrate to the center and maintain intense combustion there. The slope or angle of the bosh-wall must be such as to give proper support to the charge, which rests upon it, and yet allow the solid
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Gas Cleaning - 296
coke to slip down; aD angle of 80° is preferred. The height is limited to the height of the smelting zone. These conditions limit the diameter of the bosh to 22 ft. From the top of the bosh the stack wall must decrease in diameter to the throat to give room for the descending chai^ to swell by reactions that occur in its downward prepress. This leaves, at the throat, a diameter suitable for the proper distribution of charge. Furnaces have been built higher than 100 ft.,- but such height has been found to be excessive, especially for fine ores; and the best practice calls for 90 ft. or leas.
OAS CLEANING . The top gas coming away from a blast-furuace, especially when smelting fine ore, carries much dust caused by the agitation of the blast. Some of this is settled out in the dxist-catcher, but the gas still remains quite dusty. When the gas is subsequently burned at the stoves the dust settles in the checker work and at the boilers it attaches itself to the stoves. If the gas is cleaned it bums more efficiently Mid, moreover, it can then be used for driving a gas-engine blower plant.
Fig. 157 gives the views of a scrubber plant for gas cleaning for stoves and boilers for two furnaces. Of the figure, Tt^, is a front elevation, X, a side elevation, Y, a separate elevation of the dust-catcher, and Z a plan view of one of the scrubbers, to show the arrangement of the water sprays. The gases from the duatKatchers of the two furnaces are imited in the 7-ft. gas main, a, to go to either of two dust-catchers, 6, each leading to the scrubbers, a and s'. It should be here noticed that one of the scrubbers, if pushed, will clean 40,000 cu. ft. of gas per minute while the other may be by-passed and used as a spare, though for the best work both are used. In each scrubbing tower, 14 ft. diameter by 74 ft. high, are two sets of ring pipes each sending up jets of water, these effectively cleaning the ascending gas, wetting down the dust particles and causing their fall to the sump below. Here is a goose-neck siphon that permits the discharge of the wetted particles or ore pulp. The cleaned gas from the top of the towers passes hy the down-comer, c, c, to dust catchers d, d, that dischai^ into the cleaD-^as main E, for use at the stoves and boilers.
The Hot-Blast Stoves
The efficient operation of an iron blast-furnace requires that the-air entering at the tuyeres be brought, generally to a red heat (500 to 750° C). To do this the furnace is equipped with three or four (as in Fig. 152) regenerative firebrick stoves 80 ft. high and 14 ft. diameter. The Cowper stove (of Fig. 158), for example, consists of a tight shell, like a boiler shell of steel plates, lined with firebrick, and containing a
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Iron Ores And Their Smelting
checker-work of bricks of special shape laid in open order so as to hav-e numerous openings or passages from the top to the bottom of the stove.
The gas from the furnace, containing 24 per cent CO, which in burning supplies the heat, flowing along the underground flue from the gooseneck before mentioned, enter the stove g, while the air is admitted at a.
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The Hot-Blast Stove 297
the two mingling and burning in the vBrttcaJ circular flue /, and heating the checker-work in their passage to the valve s, as shown by the arrows. The gas thence passes by an underground flue to a tall stack, 200 ft. high, shown behind the stoves in Fig. 152. Gas having burned in this way a half hour, the stove becomes heated. The valves g, a, and « are closed, and the cold-air valve at c near the bottom of the stove is opened, admitting air under pressure from the blowing engine ; the hot-air valve h, at the bottom of the flue /, being at the same time opened. The air rises through the hot checker-work, descends the flue /, and passes out at <j to the brick-Uned hot-blast main, and to the furnace.
Meanwhile the gas has been turned into the other stoves, and b heating them in the same way. After the blast has been received in the first stove a half hour it is turned into the next heated stove, and so on. The extensive surface of the checker-work serves to absorb a large amount of heat from the buming-gas, and to impart the heat subsequently to the blast-air. It will be noticed that the air epters at the coldest and leaves at the hottest part of the stove. It flows in a direction opposite to that cf the burning gas, thus insuring the maximum rise in temperature. In the course of half an hour, the
hot air, leaving the stove, falls at y,^ i58.— Cowper Hot-blwt Stove, least 100° C. in temperatiue.
Not all the ga? from the furnace is needed for heatii^ the stoves, and ft
Iron Ores And Their Smelting
poitioD is burned under the boilers of the plant for making steam for power. The amount of steam thus available is suffident to run the Mowing-en^nes, hcusting-caigiiies, hoisting-mechanism, and all machinery belonging to the furnace. At some plants the surplus gas, after the stoves have been sup- {died, has been cleaned from dust and used in gas-engmes. Power can be gamed in this way and is more available for rolling mill or other
Formerly the air was heated in irtm-iupe stoves, the air circulating through a nest of pipes inclosed in a furnace or brick heating-chamber. Hieae are no longer used, but have been supplanted by the tcgenerative stoves just described.
BLAST-FURITACE AND ACCESSORIES F^. 159 is a plan <rf the Uaat-furoaces, showing the course of the gases throughout the plant. From near the top of the bUst-fumace, as well
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Flo. 159.— Plan of Fi
shown in Fig. 152, come away the two branches of the down-comer which uniting in one enter the top of the dust-catcher. This settles out much cJ the flue-dust in the furnace gases produced by the heavy blast in the fui^ nace. The rest is removed at the gas washer, so that a dust-free gas may be used at the boilers and the stoves. A gas-main, taken off at the gas-washer, passes along the front of the boilers with branches to each, where it is burned for the generation of the steam needed for driving the blowing
The Blowing Engine 290
engine, that taking some 9 per cent of the total beat. The other branch from the gas-washer paaaes along In front of the stoves, some 14 per cent of the total heat being there utilized. From the blowing-engine house, where there are three blowing engines, cornea away the cold-blaat main to the stoves. Here the air is heated 500° to 750° C, wid goes by the hot-blast main to the bustle pipe, thence through the tuyeres, into the blast-
FiQ. 160, — Blowing-engine.
furnace. Both the hot-blast main and the bustle pipe are lined with tilb to prevent heat loss by radiation from the pipes.
Blowing Engines. — Fig. 160 gives a view of a vertical blowing-engine, having an aiiM^linder at the top 7 ft. in diameter and of a 5-ft. stroke. The cylinder displaces 385 cu. ft. air per revolution, or 15,400 cu. ft. free air per minute, and deUvers it at 15-lb. pressure. The air-admission and discharge valves are arranged to operate positively, and to open and
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300 Iron Ores And Their Smelting
close at exaotly the right moment. Enough of these engmes are uistaHed to supply the amount of air required by the furnace.
Operation Of The Blast-Furnace
The ore is dumped into the furnace with 45 to 60 per cent of ita wei^t of coke, and the limestone needed to form the predetermined slag. The furnace should be at least 65 ft. high, and is now built 80 to 100 ft. hi^. It is kept full of stock, and the combustion of the coke is supported l^ the air introduced at high pressure at the tuyeres. As smelting progresses, the coke burns, the slag and iron produced from the charge is withdrawn, and the surface or stock-line sinks. Thus the remov^ of molten products below and the addition of fresh stock above, cause the greatest producticm oi beat next the tuyeres, where the coke largely bums, the temperature decreasing toward the stock-line. The -actual melting zone, or zone of fusion, extends upward through the bosh region. The most intense combustion occurs within 4 ft. of the tuyeres, where an excess of air, driven in under hi|^ pressure, burns the coke to carbon dioxide. In the reaction, the COs may be said to dissolve the carbon, it being as follows:
The reaction being endothermic, l^aens the temperature in this seccmd re^OD. The temperature is so high, however, that the glowing coke reduces to iron any iron oxide that descends as far down in the furnace as this re^on. The rifong gas consists of COCOa and N. The carbon monoxide, in the ascent reduces the iron oxide to iron. On reaching the lower part <rf the furnace the iron, with carbon taken from the CO, forma pig iron. The iron takes up silica, phosphorus, and sulphur from the earthy constituents of the chai^. The carbon amounts to 3 or 4 per cent, and the other impurities to 2 per cent the we^ht of the product. It is the 5 per cent of the metalloids present in the pig that makes it fusible.
We have, therefore, in the blast-furnace beginning from above, three zones:
(1) The zone of preparation, where CO2 is driven from the limestone and moisture from the charge.
(2) The zone of reduction, where the CO of the rising gas reduces the iron ore, first to the ferrous form, then to iron, and where the iron in a spongy or open form absorbs carbon from the reducing gas.
(3) The zone (A fusion, where the temperature of the furnace is lu^ and the slag is formed, the iron at the same time absorbing sihcon, phosphorus, and sulpur.
In the upper zone of the furnace, the carbon dioxide of the limeet«ue is expelled, leaving quicklime (CaO) ready for fluxing the gangue or waste
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Blast-Furnace Operation 301
matter of the charge. The action is endothennic, leBaening the heat of the escaping gases. The quantity of limestone needed to form a suitable slag is calculated in advance.
The gas varies in composition, but commonly contains 61 per cent nitrogen, from 10 to 17 per cent CO2, and 22 to 27 per cent CO. The carbon monoxide is the combustible constituent of gas that produces the heat when the gas is burned in the stoves and boilers.
In the foregoing description it has been assumed that coke is the fuel employed as is true in most cases. Anthracite coal has been xised when cheap enough to compete with coke, but even then a more satififactory result is obtained when coke forms part of the charge. Furnaces in which a part of the fuel is anthracite are called anthracite furnaces, but the name ia somewhat misleading. Other furnaces use charcoal exclusively. Charcoal is supposed to given an iron of great toughness that is particularly valuable for cast-iron car-wheels and other castings requiring touf^ess. Its superiority over other kinds having the same ccnstituticai has been widely disputed, but there is testimony in favor of charcoal-iron.
Thirty years ago blast-furnace practice was regulated by rule-of-thumb methods. They were " bom of a bigoted belief, on the part of ignorant fumace-men, that particular ores and fuel could be worked in a furnace only on special lines, and that it was impious to drive a fmnace faster than a certain rate estabhshed by time-worn tradition." In 1879 certain experiments made at the Edgar Thompson Steel Works, Pittsburg, Pa., showed that it was possible to increase the output of a furnace enormously by increasing the air-supply. It was also found that the amount of air, not the pressure, determined the rapidity. Under the new system it was thought necessary to make a steep-angle bosh (80°) resembling that in Fig. 155 more than that in Fig. IM. With the more rapid driving, reduction decreased, and the slag contained more iron. To secure the reduction, the fuel had to be kept high, using one ton of coke per ton of pig iron produced, and where coke was expensive this was a serious matter. E. C. Potter at the Illinois Steel Works, South Chicago, showed that by reducii^ the bosh-angle to 75° and using somewhat less blast, it was possible to cut the coke consumption from 2240 lb. to 1800, or even 1750, per ton of p^ produced. Furnaces with large hearths were then built, which also increased capacity.
Blowing-in. — The furnace is first dried several days by a wood fire in the crucible. The lower part, h^way up the bosh, is filled with cord-wood. Upon this is placed a heavy bed of coke with limestone to flux the coke-ash, followed by successive layers of the normal charge of coke, with gradually increasing amounts of ore and limestone, and decreasing quantities of slag, \mtil the normal charge of ore and flux is reached. The wood is ignited at the tuyeres, and a weak blast of air supplied. The pressure is gradually
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302 Iron Ores And Their Smelting
increaaed during twenty-four hours, and the furnace becoming entirely filled during this time the regular pressure ia reached.
Regular Operation. — ^The old way of chaining the furnace by hand is as follows: Ores, flux, and fuel (the " stock ") is brought in buggies frcsn the stock-house to the scales, weighed, hoisted to the top of the tutnace, ex timnel-head, wheeled by top-fillers to the bells, and dumped evenly, the fuel separately, the ore and " stone " (limestone) together. The furnace ia kept full to a level just below the outlet <A the down-comer. In automatic chaining, as indicated in Fig. 154, the stock is charged to the skips, as has been described, and is hoisted to the fumace-top and dumped into the double hopper. No top-fillers are needed, and the bells are often operated from the ground-level, so that no attendant is needed at the tunnel-head.
nCREOOUUUTIES m BLAST-FDIUIACB OPERATION
In the upper part of the smelting zone semi-fuaed material may attach itself to the walls of the furnace, and this is the be^nning of an accretion which fonns a " scaffold." This scaffold or hanging may gradually build out imtil it arches over, holding up the charge and nearly stopping the furnace. Such a miahap is more liable to occur with fine ore such as that of the Mesabi range. Sometimes this scaffold may be broken down by suddenly cutting off the blast pressure, and allowing the full weight o£ the charge to come upon the obstruction. If this proves ineffective, then, by cutting a hole through the wall of the furnace, the obstruction may be melted out by the aid of a blow-pipe burning oil or gas.
Sometimes the scaffolding may give way in part, causing slips by which material is suddenly precipitated bo the hearth, and there lesults an upward rush of gases resembling an explosion. This may do damage to the charging, interrupt operations and throw stock out of the top of the furnace. Scnne furnaces are provided with explosion-doois or valves which open under the sudden pressure and relieve the strain.
So much cold material, precipitated toward the hearth by a slip, tends also to caxise a " freezing " or solidification of the slagged material near or over the tuyeres. Th^ 8(Jid layer may sometimes be broken away by driving in a steel bar to enable the blast again to enter. This may lesult in the heating up at this point, and a final melting away of the obetructi<»i. Sometimes it is necessary to melt through the frozen material by the aid <rf a blow-pipe, or, in extreme cases, to break through with the ud oi explosives.
Or the molten iron near the metal notch may get so cold as to solidify so that it becomes impossible to enter. Then another tap-hole must be made by boring through into the crucible at a higher level. When the furnace is again regularly working the heat gradually descends until the
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Disposal Of Slag 303
whole contents of the hearth are melted out, when the regular tap-hole can be again tised.
Irregularities in the mnelting change the character of the iron made. Thus cold material coming down to the hearth will chill the smelting z«ne, and cause the silica to be low and the sulphur high, that is, will make white iron when gray or soft iron is desired.
Disposal Of Slag Or Cihdbr
On account of its low specific gravity, slag floats upon the iron. The iron occupies the lower part of the crucible, and accumulates until it reaches
Flo. 161. — Slag Ladle aod Locomotive.
the tuyeres, when it should be drawn oft. Every Ij to two hours, the plumed cinder-note is pierced with a pointed steel rod, and the cinder above the level allowed to Sow out. It flows along a cast-iron launder a distance of 15 to 30 ft. and falls into a 14-ton slc^-car standing on a track below. When loaded the car is hauled by a locomotive to the dump that may be a mile away, and the contents of the ladle is poured out at the side of the track. The track is gradually raised and moved outward toward the edge of the dump as it grows.
A lai^ coke furnace, yielding 500 tons of pig daily, when smelting ores of good grade, produces 300 tons of cinder. In smelting sihcious ores, the quantity of slag may be twice as great.
Iron Ores And Their Smelting
DISPOSAL OF PIO mOH
Every four to six hours the metat is tapped from the furnace, 50 tons at a time. The flow is started by a pointed steel bar which is driven through the clay-plugged iron-notch or tap-hole. A clay-lined launder conducts the flow to ladles similar to the cinder car. Distant from the furnace 10 to 15 ft. is a cross-channel made in the sand. The slag that floats on the stream of iron is diverted into the cross-channel with a skimmer. An iron plate is placed across the Hewing stream in such a way as to permit the heavy iron to flow beneath, while the light slag is diverted. A charcoal furnace, having an <mtput of 100 tons per day, produces so little cinder that none is tapped at the cinder-nobcb, but flowing out with the metal it is akinuned as above described. It is run outside the cast-house upon the ground, is allowed to cool, and is then broken up and carted away. The iron contained in the ladles ia called " direct metal " and may be taken to the steel works'and used in molten form.
In practice elsewhere, the iron is cast in molds in the sand of the floor
Fio. 162,— Heyl A Patttraon Pig-cssting Machino.
of the castrhouse. The floor, 40 ft. wide and 80 to 150 ft. long, consiBts of the sand in which the depressions are molded and connected by a main chaimel or runner, which receives the molten iron flowing from the furnace. The molds fill successively and form pigs of iron weighing 150 lb. each.
To reduce the cost of handling the pig metal, and to pve a product anooth and free from adhering sand, casting machines have been mtroduced in modem plants. Fig. 162 is a Heyl & Patterson pig-casting machine, conasting of an endless-chain conveyor composed of a series of molds, each capable of holding 120 lb: iron. The iron, brought from the furnace in a large ladle shown at d, Fig. 162, is poured into the molds as they travel slowly along. The pig-iron chiUs quickly, and by the time it reaches the discharge end, it consists of solid pigs of iron and drops into the railroad car that is placed in position to receive it. The molds on their return, inverted, take at c a spray of whitewash, the water of which quickly dries by the heat of the mold. It leaves a coating of lime inside that prevents the iron from adhering. Mechanical casting has the advantage over casting in sand-molds that it does away with the hot and severe work of
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Dry-Air Blast
breakmg and handling the be bome, and there always is Blowing-down. — When a stopped, and a layer of coke blast the stock-line deacenda and cinder can be tapped Finally the blast is stopped, drawn through a hole broken
In hot weather the work can hardly difficulty in getting or keeping the men. furnace is to be put out of blast, charging is s added for the last charge. With continued and the opeTation prepresses as long as iron out, the blast being gradually diminished, and the remainder of the contents is wil^- in the brickwork near the bottom.
Drt-Air Blast
This aims to regulate the one big variable in the operation of the blast-furnace. All the other materials, the ore, stone, and cc^ vary at the most but a few per cent in their compo«tion from time to time, while the moisture in the air may vary more than 100 per cent from day to day, and of this air the production of a ton of iron requires almost double the weight of the other raw materials.
The moisture in the atmosphere is removed by refrigeration previous to the introduction of air into the stoves. Any moisture, on entering the furnace at the temperature at the tuyeres, is at once dissociated by the intense heat into the component gases according to the reaction.
or per poimd of hydrogen 29,- 030 pound-calories per pound. The amoimt of air necessary to bum the fuel for smelting 100 lb. of pig varies with the temperature of the blast, but a fair average may be taken at 5300 cu. ft. With the amount of moisture contained under average conditions, as pre-by
306 Iron Ores And Their Smelting
viously assumed for Pennsylvania (3.44 grains) the total moistuie will then be:
1545 calories X2. 6 ="4020 calories.
The above figure represents the heat lost per hundred pounds at iron smelted, due to the moisture in the atmosphere or, expressed in coke consumption, somewhat over 50 lb. of coke per ton of pig made.
The Gayley process, as installed at several plants, has removed by cooling to 25° or 30° F. approximately 65 to 70 lb. of water per ton of pig smelted during some of the more hmnid months of the year, a theoretical saving in the consumption of coke of some 55 lb. per ton of iron, but the actual saving has proven far greater.
It will be seen that the actual saving far exceeds any that can be theoretically accounted for, either by the elimination of moisture or by the rise of temperature in the blast.
There has been much discussion regarding it, but probably the greatest saving is in reality attributable to the securing of imiformly favorable operating conditions. R^ularity is a prime essential for economical running and with these conditions \miformly good there is no need for the excess fuel necessary to provide for contingencies.
It would not, however, be the part of widsom to assume that such enormous saving could be made at all plants and under all conditions, but it is probably safe to say that the average plant can decrease ite coke consumption at least 12 per cent and increase the production 10 per cent, while in very many cases it is perfectly feasible to raise these percentages to 14 per cent and 12 per cent respectively. This, of course, refers to what may be relied upon the year through and not merely for short periods under the stress of record breaking output.
Chbhical Reactions Of The Blast-Fokitacb
A blast-furnace may be likened to an immense gas-producer in which there is a column, 70 ft. high, of alternate layers of coke, iron ore, and flux. The column ranges in temperature from a heat that shows no color at the tiiroat, to a white heat at the tuyeres.
The hot air of the blast, entering at the tuyeres, strikes the white-hot coke with the immediate formation of CO2 followed by an instantaneous reduction to CO. The air therefore need only bum the fuel to CO as indicated by the following reaction :
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Reactions Of The Blast-Furnace 307
Per pound of carbon burned 2415 pound-calories are generated. Since
23 per cent erf air is oxygen, and at the sea-level 1 lb. of air equals 12.38
16 12 30 cu. ft., we have — X ' —72 cu. ft. of air per pound of carbon, or 61
cu. ft. per pound of coke of 85 per cent carbon.
Fig. 164 shows graphically the chemical reactions under a set of conditions assumed, while the temperature and places where the reactions take place are shown in the section of the furnace at the extreme left in the diagram. To produce a ton of pig iron (2240 lb.) there is to be used 3520 lb.
P'i<i. 164. — ChemicSil Reactions of the Blast-Funmce.
of 60 per cent iron ore containing 3020 lb. FeaOa, 1888 lb, coke, ^nd lOlO lb. limestone.
At the txmnel-head, the iron ore (FezOa) plunges into an atmosphere of 24 per cent CO, 16 per cent COa, and 60 per cent of N at a temperature of 260° C. Reduction of ferric oxide to FeaO* by the CO begins thus:
(4) SFeaOs + CO = 2Fe304 + CO3 3X199,400 29,000 2X270,800 97,000= +11,400 Cal.
The reaction is completed at a temperature of 450" C. when the ore has reached a depth of 10 ft., shown at 2 of the diagram. During this period the peculiar reaction resulting in carbon-deposition begins, caused by the reaction of the gas on the ore, forming a deposit of soot or carbon in the pores.
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308 Iron Ores And Their Smelting
CoDtinumg tlie descent the ore undergoes further reduction. At a depth of 19 ft. and a temperature of 600° C, the FesOi formed, as shown above, has become further reduced to FeO, as indicated in column 4, the reaction being as follows:
265,800 29,000 3X66,400 97,000-= +1400 Cal.
The FeO thus formed, impregnated with carbon (see column 7) descends with little chuige, until at a depth of 26 ft. and at a temperature of 700° C, the CO of the gas reacts upon it, and spongy iron begins to form. The reaction is complete at 800° C, and at a depth of 32 ft., and is as follows:
In the passage downward the limestone gradually loses its CO2, and at thU point the expulsion is complete, as indicated at column 8. The quicklime, column 9, thus formed, unites at the zone of fusion and fluxes the sUica of the charge. From the depth of 19 ft. to the depth at which all carbon dioxide is expelled, that is between the temperatures 550° and 880° C, the CO2 reacts upon coke, dissolving it according to the following reaction:
Thus heat is absorbed from the gas, and some coke is consumed. The coke, however, as can be seen from column 6, remains but little changed until it reaches the region of the tuyeres.
Below the 32-ft, level at 800° C, reactions practically cease, the chief action now being a reduction of a small amount of FeO, left undecomposed by the CO. This is gradually reduced (see column 4) by the glowing coke as follows:
Silicon having less affinity for oxygen than carbon at a high temperature is formed from the reduction of the silica, and as a metalloid enters the pig iron after the following reaction:
Si02+2C = 2C0+Si.
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Heat Balance Op The Blast-Furnace 309
Below the 32-ft. level, the temperature rises gradually and uniformly until the inteiwe combustion at the tuyeres produce 1500° C. as a
Of the air entering the furnace, 77 per cent ia nitrogen, and of the escaping gas 60 per cent, thus showing nitrogen to be by far the largest constituent present. As is shown in column 12, nearly three tons of nitrogen pass through the furnace for each ton of pig iron produced. At the hi^ temperature of the lower part of the furnace, potassium cyanide is formed, the potash of the coke-ash uniting with carbon and nitrogen to form the Bait. It decomposes before the top of the charge is reached.
Referring to columns 10 and 11, we note that the COs formed so freely at the tuyeres, is at once (column 10) changed to CO. The carbon monoxide rises unchanged until it reaches the 32-ft. level, when it begins to act on the iron oxides with the formation of COz. The carbon dioxide from this source united with that from the limestone ia the total of the escaping CDs gas.
Sulphur occurring as FeS in the coke and as pyrite in the ore, is speedily driven off by the heat of the furnace, giving FeS. The sulphur of the FeS is taken up by quicklime, and enters the slag as calciimi sulphide according to the following reaction:
Thus it is separated from the iron, upon which It would have an Injurioxis effect.
The Heat Balance Op The Blast-P1Tr1Tacb
The heat yielded by the fuel and blast on the one hand, and that absorbed in the various reactions, taken by the blast, and lost by radiation, may be stated for a particular case as follows:
To produce a ton (2240 lb.) of pig iron of the composition, carbon 42 per cent, siHcon 1.35 per cent, manganese 0.64 per cent and iron 93.6 per cent, there was needed 4093 lb. of ore, 795 lb. of limestone, and 1682 lb. of coke, and besides the 2240 lb. of pig iron, there was yielded 1010 lb. of slag and 89 lb. of flue-dust. There was blown into the furnace 6673 lb. of air (say 80,000 cu. ft.) and there came away 9309 lb. of top-gas of thecompoeitionby weight CO2, 22.3 percent; CO, 22.4 per cent; HaOCHi, 0.1 per cent and nitrogen 54.9 per cent. The coke contained 89 per cent of fixed carbon and 9.4 per cent ash, while the composition of the ore (including a Ettle scrap returned) was 51.2 per cent iron and 0.74 per cent manganese. The specific heat of air-blast was 0.248, the air carrying 5,5 grains of moisture per cubic foot, and the average blast temperature was 672" C. The heat balance sheet for 1 ton of pig iron is:
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310 Iron Ores And Their Smelting
Qmtraltd by Ctl.
CombUBtioQ of carbon to CO 2,220,000
Combustian of carbon to COi 3,752,000
Heat conteat of blast air 1,147,000
Heat conteot of moiature in air 20,000
Contumtd by
Cml.
Roduction of Fe,0, 272,000
Raduction of MnO. . , .- 27,000
Reduction of SiOi 232,000
C&lcinatioD o( carbooatea 400,000 Cal. or 5i%
Dissociatioii of moisture in the blast 220,000 Cal. or 3.1%
Carried off with the iron 635,O0OCal. or 8.9%
Carried off with the alag 55O,00OCal. or 7.1%
Carried off with the dry top-gaa 417,000 Cal. or 5.9%
Carried off with the moisture in the top-gas 388,000 Cal. or 5.4%
Radiation, cooling-water and unaccounted for 631,000 Cal. or 8.8%
Total calories 7.179,000 Cal. or 100,0%
It should also be noted that the heat carried off by 1 lb. of iron is 261 calories, and that by 1 lb. of slag 483 calories. By a careful study of the above figures one may arrive at a just estimate of the value of the various furnace operations, and so can compare them .with the peiformance of other furnaces.
BtTRDEiniTG THE BLAST FURHACB
This involves the calculation of the proper proportion of ore and ftux needed for the production in the furnace of a slag of suitable c<HQpo6ition. In order to accomplish this we must have an analysis of the materials of the charge and of the fuel. The furnace must work freely and regularly, and must produce the kind of iron desired as more particxUarly shown on page 3 15. This is accomplished by so burdening the furnace as to produce a slag of the proper composition, and by property regulating the qtiaotity of fuel and the temperature of the blast.
The Slag. — ^This results from the melting tc^ether of the non-volatile solid constituents of the charge, that is, the silica and bases of the ore and fluxes, since slags are essentially silicates of these bases.
Below we give the composition of typical slags that have proved altogether satisfactory in practice.
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Charge Calculation For The Blast-Furnace
Slaq.
Iboh.
SiO,
MtO, CO
Bi
a
AvengM for Hot Fumaeefl.
Cuban ore
Spanish ore
tr. 0.02 0.02
Avenges Tor Cool Furnaces
Cuban ore
Spanish ore
In these flags the ratio of 3iOz to CaO+MgO will average aa 33 to 63 or ae 1 to 1 .6. Alumina is not regarded as an acid or a base, but as a neutral constituent disaolving in the slag. As eeen in the table a hot-running furnace reduces mote silicon to enter the pig, producing gray or soft or foundry pig; when run cool more like the gray foi^, the mottled or white iitm; also the retained sulphur in Uie pig is higher.
The weight of the fuel of chai^ for a good-eised furnace should be such . as will fill a skip of 6 long tons capacity, the iron ore and " stone " (limestone) being separately hoisted and put in to the furnace. These two skip loads are called a " round." About a ton of coke is needed to produce a ton of pig, thou^ for a cool furnace as little as 1600 to 1800 lb. of coke haa been used. Twotonsof iron ore of 50 per cent Fe^ould yield one ton of pig. We may decide, that, as experience suggests, we need one-fourth of its weight of limestone. On this basis we will prepare the charge sheet as shown on page 312.
The amount at the items are written in, with their percentage composition. The corresponding weights of the elements afe calculated through and their totals obtained. The 11,800 lb. of iron is to have 1.5 per cent its wei^t of silicon, or 1751b. Now 1 .5 per cent Si corresponds to fjof silica, or 374 lb., and this subtracted from the total, leaves 1589 lb. silica for the slag. This multiplied by 1 .6, as already explained, g^ves us 2534 lb. Subtracted from the total it shows an excess of 672 lb., or of limestone twice that, say 1300 lb., so only 4200 lb. is needed. Erase where needed, put in the new figure of 4200 lb., recalculate and obtain new results, which should be neariy correct. If not, a second correction can be carried out.
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Iron Ores And Their Smeltinq Charge Sheet
Weifbt.
aio..
Ft
ClO-MtO.
P.
<£?..
Po„„..
s;.
Pound..
p»«,.
Pumdr
Gogebic iron ore...
i.a
For DM
_
-needed
C 3,5
Where manganese exists in the ore, about one-third of it enters the slag as MnO, making it more fluid. The rest accompanies the pig iron. A little iron may, as FeO, enter the slag, but this loss to the pig iron is more than made up by the additions of carbon and silicon that it receives.
GEHERAL ARKAIfGEHENT OF THE BLAST-FtTRJNACE PLANT
Fig. 165 shows the general arrangement of a two-fumace plant. All parts of the plant are reached by raihxiad tracks on short easy curves. The ore is brought in by ore-cargo steamers and stored for the winter period in an extensive ore storage yard. From the yard it is reclaimed by a traveling crane and stored in ore pockets adjoining the furnaces and next to the coke pockets. The coke is brought in by an overhead track and imloaded to the pockets. The blast-furnaces, marked respectively A and B, have each their cast^house where the iron can be cast in sand-beds. At the end of the building is the pig-breaker where the pig is broken in order to determine by fracture ite grade. Between the furnaces are seen the stoves, four for each furnace. The large building, between the cast houses, A and B, may be called the power house, and contains the engines, with their blowers and the boilers heated by fumace^as, also the machine shop. To the right of the yard is the building for the pig-castii^ machine. The molten pig iron from the furnaces is brought by ladle car to the " ladle house " adjoining the pig-casting house and there poured into the molds of the machine. Alongside each cast-house is the hot cinder track where the cinder-car is run in to receive the molten slag as it is tapped from the furnace, and thence taken to some distant dump.
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Iron Blast-Furnace Plant
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Iron Ores And Their Smelting
PIG IROn
Tbe inm produced in the blast-furnace is not pure, but contains 3} to 4 per cmt carbon and 1} to 3 per cent silicon. Some of the carbon is combined chemically, some separated as graphite. If a large proportion ia combined, the metal is hard and the fracture of the iron looks white. If a large proportion is free, the fracture is gray or black with scales of graphite, and the iron is soft and tough.
Under " Chemical Reactions of the Iron Blastr-fumace," Equation (3) indicates the formation of carbon, and Equation (8) the reduction of silica to silicon, both elements entering the jug iron. A sm^ amount of sulphur, seldom less than 0.2 and <rften 0.26 per cent or more, is present. As the amount increases above 0.1 per cent the iron becomes harder and more brittle.
The percentage of silicon and sulphur in the iron depends in large measxae upon furnace-conditions; hence it can be controlled; but all the phosphorus present enters tbe pig iron. In pig iron for steel manufacture by the usual, or acid Bessemer process, the phosphorus in the pig must not exceed 0.10 per cent. Therefore, in the ore, it must not be higher than 0.05 or 0.06 per cent. In the acid Bessemer process the phosphorus is not eliminated, and it tends to make steel red-short (brittle when hot) a quality that interferes with the subsequent rolling. Phosi^ionic, on the other hand, imparts tbe quality of fluidity to cast-iron. Iron that contains 3 per cent P is in demand where intricate castings are to be made, and can be used where brittleness is (rf minor importance.
Cast iron as compared with steel and wrought iron has the following characteristice:
(1) It is brittle because of the presence of the metalloids, carbon and silicon.
(2) Becauseof the presence of the metalloids it is fusible; anditderives thus the most valuable property. It runs freely from the blast-furnace and can be cast in intricate molds to form castings of any kind. Wrought iron at the same temperature would be pasty and would not run. Steel, which is intermediate between wrought iron and cast iron in the contained carbm, can be made into castings, however, but not readily like cast inm. The making of steel castings is becoming more common.
(3) It cannot be forged either hot or cold.
Classification Op Pig Iron
P^ iron for further treatment or use may be thus distinguidied: Mill Iron, — For puddling, a pig low in silicon is needed, but otherwise of a different quality.
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Classifications Of Pig Iron
' Pig. — By this is meant an iron containing lees than 0.10 per cent phosphorus and less than 0.05 per cent sulphur.
Basic Iron. — ^Thia should be low in siUcon, that has been cast in an endless-chain casting-machine, thus being fiee from sand. Silica attacks the lining of a basic-lined open-hearth.
Halleable-wm Pig. — Used in making malleable iron castings. It is n<m-BeBaemer, low In hHcod and graphildc carbon.
Charcoal Iron. — This is made in charcoal furnaces, and is used for special purposes in the foundry — as, for example, in m airing car wheels.
Foundry I^. — Is used for making castings for all purposes. The iron ^ould leadily fill the mold and not shrink much when cast. Otherwise a grade of itaa is used to suit the purposes to which the casting is to be put.
Grading Pig-iron by Fracture. — The pigs are broken in two, either over a wedge-shaped block or in a machine, and the fracture is observed. Foundry No. 1 is dark gray in color, the grain large and even; foundry No. 2 has a small uneven grain and ia lighter in color; foundry No. 3 is cloee^rained and light in color but has less than 3 per cent silicon, in fact
Grading by Anafysis. — This is a more reliable method than by fracture. The character of the grades of Alabama pig iron is indicated by the table below:
Alabama Pio Iroit
Silver gray . . . No. 1 Btrft. . . . No. 2 soft.... No. 1 foimdiy. No. 2 foundry. No. 3 foundry.
Gray forge
Mottled
White
S 53 3.40 3 55
Tliis table shows the mcrease of comlnned carbon, and the decrease of silicon, as the grade approaches white iron.
The first grades are more difficult to make, and command a higher iwice.
Pittsburg Pig Iron.— For the Pittsburg district we give a similar table but with the silicon, on the whole, lower and disregarding the carb<Hi.
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Iron Ores And Their Smelting
Gr»y forge
Basic (chill cost) . Strang foundry and oar-wheel
Low phoaphonu. No. 2 foundry. , .
Lien than 1,00
0.75 to 1.60 I 00 to 2.00 I.00to2.00 1.75 to 2,25
Over 0.05 0.05 and lees
0.05 and leas 0.05 and leas 0.035 and lest 0.05 and Ima
0.03 and less 0.10 and leas 0.035 and leM over 1.00
Carbon. — This occurs graphitic and combined, in ordinary pig up to 4.5 per cent, and in high manganese and chrome iron to as much as 7 per cent. When molten the carbon is regarded as being combined, but in cooling more or less separates as graphitic carbon. When much of the latter separates the fracture is darker in color and softer than where the carbon remains combined; it is well suited to machining, though not so strong as when the carbon is combined. In smelting, to obtain a pig that will be high in graphite, the temperature of the blast and the proper-tion of fuel should be high, so as to secure a good reduction, and that much carbon shall be taken up from the fuel. This also ensures reduction of much silicon to enter the pig, and cooling compels carbon to take the graphitic fonn.
SiliGOil. — This is reduced from the silica of the change at the hearth of the blast^-fumace. It then dissolves in the forming iron. For a high-silicon iron a high temperature is here needed and this is accomplished by a light burden and a good hot blast. Iron, containing as much as 20 per cent ailicoii has been made in the blast-furnace, and when the pig contains more than 6 per cent it is called ferro-nlicon, a product much vsed in steel making.
Phosphorus. — All of this element present in the ore is readily reduced in the blast-furnace to a phosphide which combines with the iron. It makes the iron more fluid so that it better fills the mold but the casting 18 more brittle.
Manganese. — This is reduced like the iron. It aids the pig in holding the carbon in combined form. Manganese increases the strength and fluidity of the pig, and makes it harder and leas fusible. Likewise it tends to remove oxygen and sulphur from the iron, and to counteract the detrimental effect of other impurities. To make ferro-manganese alloys, much used in st«e1-maldng, a separate blasts-furnace is operated, using a very hot blast and a light burden (a high fuel) , since manganese is difficult to reduce. An alloy containing 10 to 25 per cent manganese b called spiegeleisen
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Characteristics Op Pig Iron 317
(imrror-iron) because of its Bhining cryBtalline appeBrance; and when contaiitxQg 25 to 95 per cent manganese it is known as ferro-manganese. Su^hnr. — ^Thia detrimental element occuia dissolved in the pig metal as FeS. It makes it hard and brittle and tends to keep the carbon in combined form. A high, percent^e of sulphur makes porous castings, but the iron is more fluid when cast. In the blast-furnace, using a high lime slag, this tends to take it away from the iron.
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CHAPTER XXVI WROUGHT mON AND STEEL
Cast iron, becaxise of the large proportion of contained metalloids, ' which indeed makes it fusible, and easily cast, is too weak and brittle for many structural purposes:
Therefore three-fourths or more of the pig iron in the United States, together with much steel scrap, is made into steel because of its superiority as an engineering material. About 3 per cent of our pig iron is made into wrought iron, a product superior to steel for certain purposes, because of its welding quahty and ductibility as compared with ordinary Bessemer or open-hearth steel. For most engineering purposes steel is, however, superior to wrought iron and as cheap.
THE HAnUFACTUKE OF WROUGHT IRON BY THE PUDDLING PROCESS
Almost all wrought iron manufactured in the United States, about 1^ million tons per annum, is made from pig iron by the puddling process, invented in England by Henry Cort about 1780, and greatly improved by Joseph Hall, fifty years later. The grade of pig used is either gray for^ or white — see page 315. Sulphtir should not exceed 0.10 per cent and phosphorus should preferably be less than 1.0 per cent. Pig containing as much as S, 0.35 per cent and P, 2.5 to 3.0 per cent is sometimes used, since a high phosphorus in the resultant wrought iron is not so objectionable as it would be in steel. The slag, mechanically mingled with wrou^t iron, hinders it from becoming brittle under shock, the difficulty produced . by phosphorus in steel.
The Furnace. — Fig. 166 is a longitudinal section of a puddling furnace of about 1500 lb. per charge capacity. It is a reverberatory furnace having a hearth of 7 ft. by 7 ft. in size, the grate being 2 ft. 10 by 4 ft. by 9 in, mze, and relatively large for so small a hearth, in order to obtain a hig^ furnace temperature. The hearth lining of mill cinder and iron ore suffers wear and is repaired between the heats.
Paddling. — ^The pig is chained by hand into the furnace, and is rapidly
melted down in thirty to thirty-five minutes. Iron ore or mill scale
(Fe304) is now added, this taking seven to ten minutes, and the chai^ is
thoroughly mixed and cooled to a point where the slag will begin to oxidize
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Puddling Furnace Reactions 319
impurities, especially phosphorus and sulphur. As this takes place, a light fiame begins to break through the slag-covering, due to the carbon of the pig iron reacting on the oxides of the bath thus,
Fe20a+3C = 3CO+2Fc.
Flo. 166.— l«ngiludiiial Section vt a Puddling f'uniaoe.
1
■ i
■ i
w
~
/
y
/
i
y
s,.
7
"/
i"
/
/
>-
-g>'
Flo. 167.— iSoqucnre of the ItcHctiona of the Puddling ProctWH.
The CO coming in contact with the air bums to COa with its characteristic bhie flame. As the carbon monoxide increases in volume, the chai^ becomes agitated and the " boil " is in progress. The charge swelb and the slag pouts oxit at the sht beneath the working door. This slag may amount to 12 to 25 per cent of the charge. The boil continues for
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twenty to twenty-five minutes and during the time the puddler stiis the chu^ with his long-handled rabble. Toward the end of the boil the metal begins to " come to nature," and pasty masses to form in the bath, and to show above the slag. Those masses are stirred and gathered by the T&bble until all the metal becomes pasty, when the " balling " period b^ins. The metal at this time is gathered into three or four portions, each of which are rolled up into a ball made up of many particles partly welded together. The balls are rolled up near the bridge out of the flame and in the hottest place, until the puddler is ready to draw them. They are removed one by one for squeezing, after which the hearth is repaired for the next charge.
Squeezing and Working. — The balls, weighing 125 to 180 lb. each, are withdrawn, dripping with sIe^, and are carried to the jaws of a squeezer by which moat of the slag is squeezed out and then made smaller. The squeezed balls are sent to the rolls to be rolled into bars called " Muck bar." These are cut into lengths and wired into bundles, half the bars piled croBBways of the others. These bundles are reheated in a reheating furnace to welding heat and rolled again into bars. The rerolled material is known as " merchant bar," and the effect of the second rolling is to eject more slag and to form a cross-fiber structure as the result of the cross iriling.
When rolled into strips this is called "skelp." Skelp bent into shape of tubes and butt-welded or lap-welded makes iron pipe. Steel billets are similarly rolled, forming " steel skelp " and made into steel pipe.
Following is the composition of various puddled products:
C.
Si-
8,
P-
Mn.
O.Ios 0,040 0.120
Puddled bar
Wrought iron
Usually there is more than 1.0 per cent of slag in wrought iron and less than 0.2 per cent in steel. Ordinary wrought iron is practically free from manganese, while open-hearth steel will contain 0.5 per (%nt or more, hence the greater liability to rusting of steel. Meet of the wrought iron made in the United States goes at once into commerce; a little b consumed for " crucible steel " (tool steel).
Stebl-Makikg
In the year 1918, of the 39,000,000 tons made in the United States, 47 per cent was for basic steel and 33 per cent was B^semerized. This is due largely to the fact that a pure pig iron (Bessemer pig) is needed for
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Bessemer Process Of Steel Making 321
this process, while the basic open-hearth can handle the pig fnMQ impure ores high in phosphorus, which are more plentiful. Also basic open-hearth steel has become of as good quality as Bessemer steel. The duplex process, later described, combines the speedy steel-making by the converter with the efficiency of the open hearth in purifying.
STBEL-UASmG BT THE ACID BESSEMER PROCESS
The pig iron used in the Bessemer process preferably contains 1 per cent silicon and 0.5 per cent manganese, but to make a salable steel, the
Fifl. laS.— iOO-tmi Hot-n-etal Mixer.
phosphorus should be below 0.10 per cent and the sulphur below 0.08 per cent, since neither element is removed in the converter. If the silicon is above 1 per cent the large quantity of slag produced carries away iron. If far below 1 per cent, the charge does not blow hot. When manganese is high (1.5 per cent), it makes the charge sloppy, the slag then being highly Quid and eaaly ejected during the blow.
The converters in a large plant aic supplied from several blast-furnaces, and to insure a good average pig metal, it is customary to collect the product of the several furnaces in a single tilting reverberatory furnace, or hot
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metal mixer capable of holding 300 to 1300 tons of pig metal. From ihs mixer it is drawn to the converters as needed, and a regular supply ia thus assured.
The Hot-metal Mixer. — Fig. 168 is a section of a 400-ton mixer. Like the tilting open-hearth, this is carried on rollers so that its contents can be poured into a casting ladle to go to the converters.
It is driven by two 75 H.P. electric motors which act in series during pourii^, but in parallel during the return of the mixer to its normal position. It is lined with 13) in. of firebrick face, in order to cut down radiation, with 0 in. of mcignesite firebrick.
The Converter. — The conversion is done in an upright converter, lined with sihcious material held together with fireclay. Fig. 104 represents views of a converter. It is 9 ft. diameter by 15 ft. 6 in. high and is capable of treating a charge of 20 tons of pig-metal. It is swung on tnumions, through one of which the compressed air needed in operatitm enters to the tuyeres at the bottom. The eiag made in a converter ia hi^ in silica, and has but tittle effect on the lining, so that this lasts several months. The mouths of the tuyeres at the bottom come in contact with the iron oxide formed during the blow, and hence this part of the converter lasts only twenty to twenty-five hours. Thia bottom accordingly is made so that it can be replaced by another, causing a delay of twenty miautea in changing.
Converter Lining. — For the acid process the converter is Uned to the thickness of 26 to 30 in. with ganister, that is, quartz rock mixed with some clay to bind it. In a 20-ton converter, Fig. 169, there are fourteen tuyeres of well-bumed clay, 6 in. diameter by 30 in. long, each tuyere having eight ^in. holes extending from top to botttHU for the blast-air. The tuyeres are set in place upon the bottom and the ganister is rammed around them. This work is done at the bottom house (see the general plan of duplex apd electric furnace building), where are situated the bottom-ovens and the grinding and mixing machmes for preparing the ganister.
The bottoms on transfer trucks are run into ovens 16 ft. sqxiare, where they are thoroughly dried out. These are heated by coal, using a forced draft. A bottom may last for 30 to 35 heats, or only for a smgle one. The side-lining is of ganister, especially around the nose. Bottoms are changed commonly by imbolting while the vessel is bottxnn side up, then hfting it off by crane.
Operation oi the Acid-lined Converter. — The hot converter, from which the metal of a blow has just been poured, is placed in a horizontal position and 15 tons pig iron is poured into it by means of a ladle that is brought from the mixer. When the converter is in this position no metfd can flow into the tuyeres and obstruct them. After the metal is poured in, the blast (or " wind ") is applied at the rate of 25,000 cu. ft. per minute, the con-by
GoOgIc
Converter
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verter being at this time turned to the vertical position. The blast now
blows in fine streams upward through 18 in. of molten metal. Active oxidation of the manganese and silicon results and in about four minutes
, they are oxidized by the oxygen of
the air and have become slag. The carbon now begins to oxidize to CO, and this also streams upward throu^ the metal and issues with the air from tiie mouth of the converter in a body of flame. After another six minutes the flame shortens or drops, and the operator, knowing that the carbon has been eliminated, turns the con verter into horizontal position, the Fia. 170.-MixiiiB Pan, Philips & «««* b^mg at the same time shut off.
Mcl^ren, Pittobur^, In anticipation of this, a weighed
quantity of spiegel iron or " Spiegel "
has been tapped from the spiegel-cupola, where it is kept melted, into a
Fio. 171. — Worm-geared Bottom-tap Ladle, Piq, 172. — Ingot Mould.
Pittsburg Elect. Furnace Corp.
ladle. The ladle is transferred by the traveling-crane and poured into tike converter. So great has been the heat evolved by the oxidati<m of tiie
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Acid Bessemer Converting
impurities of the pig during the ten minutes of the blow that the temperature is higher than at the start, and we have a white-hot hquid consisting of comparatively pure metal. Oxidation-products remain in the bath, and the carbon and muiganese of the charge tend to reduce these, the unused carbon being in sufficient quantity to impart the desired strength to the steel. Silicon, which also is introduced, tends to dispose of gas contained in the metal. After the speigel or " recarburizer " has been added and the reactions have ended, the steel is poured from the converter into the ladle, as shown at the left of Fig. 174, also the ladle is in position to receive the steel. This, after a short interval, is carried to a pofdtion over the ingot molds into which the steel is to be teemed or poured. The teeming-ladle. Fig. 171, is " bottom-poured," tiiat is, a tap-hole and plug are arranged in the bottom, so that when the ladle is brought over the ingot mold a stream of metal drops straight downward into it until it is filled; in this way the molds are filled successively until the ladle has been emptied.
The stopper-rod, actuated by a lever-arm outside, plugs the hole from within. The end of the rod is covered by a fireclay lining, to withstand the attack of the hot metal. The ladle is tipped by operating a hand wheel or by auxiliary hoist, to pour out the slag that remains.
Fig. 172 is the ingot-mold having lugs near the top by which the mold is picked up by crane when ready for stripping, leaving the ingot standing on the car. The handle <hi the side ia where the crane takes hold when the mold is to be laid on its side for cleaning.
The metal remains until cool, after which the molds are stripped or lifted off, leaving the ingots standing. The ingot is picked up and conveyed to a reheating furnace, and finally sent to the rolls to be formed into the shapes desired for market use. Fig. 173 illustrates grapMcaUy by curves the progress of the reactions, and the elimination
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of impurities during the blow. From it we see the rate at which the easily oxidized manganese and silicon are burned and also the carbon, which is but little acted upon until these disappear, but which after Uiey are gone oxidizes rapidly. The pig contains at the begiiimng 3.5 per cent C, 1.0 per cent Si, and 0.5 per cent Mn, all being removed. The recarburizer adds to it, as Fig. 173 indicates, 1 per cent Mn, 0.7 per cent C, and 0.15 per cent Si. The manganese is added to take from the metal the oxygen absorbed during the blow; the carbon is to give the steel the required strength and hardness, and the siUctm to dispose of the gas contained in the bath.
The Basic Bessemer Process
The converter has a basic lining of dolomite mixed with tar stamped into place. The process is used where it is demrcd to treat high phosphoiiis ores. It is little used in the United States.
A typical pig, such as is used in this practice, would contain C, 4.25 per cent; Si, 1.0 per cent; S, 0.05 per cent; P, 1.5 per cent; Mn, 0.2 per cent. This is charged, using 11 tons of the pig with the addition of 2600 to 2800 lb. of burned lime to ensure a basic slag. The blown metal would still retain C, 0.03 per cent; P, 0.07 per cent; S, 0.05 per cent; while tiie slag would contain SiO, 14 per cent; CaO, 48 to 51 per cent; MgO,2to4 per cent; P206, 17 to 19 per cent; Mno and FeO, 14 to 16 per cent. The high phosphorus content makes it a good fertilizer and it is so used.
This reverberatory furnace (Fig. 175) is used in the melting down of the materials used in the manxifactuie of steel.
The Two Processes. — There are two processes of producing steel in the open-hearth furnace, called respectively, the add and the basic. The only difference in the open-hearth furnace used is that for the acid process the hearth is lined with a sand; in the basic process with basic material such as dolomite or magnesite.
The Acid Process. — By the add process the carbon, silicon, and manganese, impurities of the molten charge, are removed, but no phosphorus or sulphur is eliminated. Hence the acid open-hearth can only treat pure ores (rf the Bessemer type low in these two latter metalloids.
The Basic Process. — On the other hand the bade process can remove from the charge when melting not only carbon, silicon, and manganese, but also sulphur and phosphorus. Thus a charge of non-Bessemer material can be used, since by the basic process it is possible to eliminate the sulphur and phosphorus below the permissible limit of 0.05 per cent sulphur and 0.095 per cent phosphorus necessary for a good grade of steel.
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Open-Hearth Furnace 327
The Opbn-Hbahth Reverberatort Forhacb
Fig. 175 is s perspective view of the front of a stationary open-hearth furnace with ita three charging doors. In 183 is a transverse section of one of the tilting kind, to be later deBcribed. The furnace is fired by producer gas, see the Hughes producer, Fig. 14.
During the past fifteen years the Bessemer process has been gradually giving way to the basic open-hearth process, due to the fact that low phosphorus ore is being exhausted. It is claimed that for most purposes open-
Fio. 174, — Converter and Mixer Building.
hearth steel is better than Bessemer, but the latter gives the most satisfactory product for tin plates, and is well suited to the manufacture of rails. An important advantage in the basic open-hearth process is that it can be used for making steel from pifr iron and ore high in phosphorus.
Fig. 176 is a half sectional plan and elevations of the furnace, having water-cooled devices designated for the better preservation of the parts exposed to high heat and corrosive action. It is basic-lined with material specified by the legend annexed.
The furnace hearth H is rectangular and open at each end o t admission of air and gas at the porta C and D respectively. The roof is of silica brick, 12 in. thick. The whole furnace is heavily ironed. The entire
bottom and hearth of the furnace is built in and supported by a pan of heavy plates, riveted together and supported on I-beams resting on piers. At the skew-backs are wat«r-cooled plates set against the I-beam buck-staves to receive the thrust of the arch. On the chai^g side are shown the five charging doors, counterbalanced and lifting vertically, large enough
to enter the charging boat umng a mechanical chai^ng-machine. Fig. 181. At the middle of the bock side is the tap-hole where slag and metal are drawn ofE.
Underground, and at one side at each end, are two checker or regenerator chambers, one for the preheating or regeneration of the air, the other for the gas. The arrangement of these is well shown in section
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Open-Hearth Furnace 329
plan and elevation, in Fig. 178. The checkers are built up of 9-in, bricks into a series of pre-heated flues through which the air and gas pass to their respective horizontal flues (ind vertical uptakes, the ports delivering to the hearth of the furnace.
The general arrangement of the furnace and its accessories is given in ■ the sectional plan, Fig. 178, particularly the passaf^es and flues that lead eventually to the stack. As indicated by the figure, gas from the main gas flue and air through an open valve at the left are traveling along their respective gas and air checker chambers and cfit^rin^ the furnace by the
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Wrought Iron And Steel
Flo. 177. — Open-hearth Fumapc (transverse section).
I
Fia. ITS. — Sectional Plan of Open-hearth li'iimacc.
Reversing Valves, Open-Hearth Furnaces 331
gas and air intakes and ports at the left. They are leaving the hearth by the uptakes on the right, passing through both of the checker chambers and thence through a three-way valve at the point marked " close," " open " to the stack.
The flow of gases having passed in this direction for fifteen minutes,
the valves are reversed and the gases pass in the opposite direction for a like time. During the fifteen minutes the flame is drawn through the checker chambers highly heating them. On reversal of the gas current the producer gas and the air, before they reach the hearth, are heated by
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the checker-work and both, thus preheated, bum at a higher tempeisttue. By successive reversals the temperature of the regenerators is raised and with it the intensity of the flame of the air'-gaa mixture. This accietiai of heat can go on to the poiBt where the brick hning begins to soften and to a temperature that readily melts the furnace charge.
Figs. 179 and 180 show a sectional elevation of the throe-wsy reversing valve, already referred to. In the upper view the producer gas is seen passing through a disk-valve D, and by the port A, to the furnace, while, at the same time the escaping gases pasrang thnn^i the checker chambers at the right and the port C, go throu^ B to the chimney or stack of Fig. ITS. The valves for the admission of air and of gas ate of the disk tgrpe.
Tapping. — At F^. 178 are to be seen the two casting pits where the ladles are set, adjoining the tap-hole and tapping-spout S. When the beat is ready to tap, a bar is driven through the tap-hole and enlai^ed by reaching through from the charging side by a long bar until the slag and metal flow out freely. The tap-hole should be carefully cleaned out after each heat, then replugged with clay.
Fuels. — The fuels employed for the open hearth are natural gas, producer gas, and oil, and of these natural gas is the best -where it can be had, as indicated by the following table :
c...,.„...
Natural G«. Percent,
■X-s.?-
The heating power of natural gas is 550 Cal., and of producer gas 150 Cal. Oil is an excellent open-hearth fuel. It can be vaporized by steam at air jet, and needs no preheating. The flame of it is, however, sharp and iB liable to cut out the roof and to over-oxidize the metal of the molten bath.
The Tilting Open-hearth Furnace. — This differs from the stationary type chiefly in that the entire furnace body may be tilted or rotated through a considerable arc, thus pouring slag or metal at any stage fA the process, frequently a great advantage.
A cross-section of such a furnace is shown in the sectional elevation of an open-hearth building. Fig, 183. This shows the furnace, in meltii^ position. This type of furnace does away with tap-hole troubles, the
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Acid Open-Hearth Process
tap-hole being above the slag-line in melting podtion; also the furnace can be readily emptied between heats and drained easily to make repairs.
Mechanical Charging. — ^At 181 is a perspective view of a charging machine. In this view is shown one of a line of trucks carrying the charging boxes. These are picked up, one at a time, by means of a charging ram, thrust through the charge-door, inverted to discharge their contents of steel scrap, then withdrawn and set once more on the trick. In this way the contents of box after box is put into the fimiace.
Fia. 1S1.— CharginK Machine. THE ACID OPSn-HEAKTH PKOCBSS
Object to be Attained. — The process aims to reduce within defined limits tiie carbon, silicon, and manganese present in the charge of scrap and iron, but leaves unchanged whatever sulphur or phosphorus thei« is.
The Charge. — This, of say 100,000 lb. weight, is made up cwnmonly of scrap steel and pig iron, the usual average being 50 to 75 per cent scrap, the remainder pig, the proportiona depending on supply and cost, and being such as to produce 6 to 10 per cent of sl^ preferably of the composition 50 per cent silica and 45 per cent of FeO and MnO together.
The following are representative percentage analyses:
PLB-irc.li,
steel 8«r«p.
st«?s'Lp,
3.00to4,00 1.00to2.00 Under 1.00 0 10 0 06
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Cranputing the silicon to silica and the manganese to MnO, the latter in the chai^ should be less than half the rallca. Silica comes also from the sand attached to ordinary {ug iron and is yielded by corromon oi the edlicious bottom.
Melting Down. — By the time the charge is melted down both manganese and silicon have been oxidized, and the resultant silica has united itself to the MnO and the little iron oxide in the stock, since the melting has been effected with a natural or even a reducing flame. In the next stage, in a hot furnace with an oxidizing flame, iron ia oxidized to enter the slag, and carbon (till then but little affected) is burned off. To aid this operation 1000 to 2000 lb. of iron ore is added in calculated proportiona and reacts with the carbon of the charge thus:
Fe20a+C = reO+CO.
Ilie iron oxide enters the slag, the CO is burned to CO3. The ore is added gradually according to the judgment of the melter, faster in a hot furnace and according to the character of the slag.
The following instructive table shows the composition of the charge both, before and after melting and of the resultant slag:
a Hun, Oil Oa*.
Endol
O.Ob 0,54 50 24 21.67 23,91 45.58
o.«
Mi':
r:.:
FeO
Hie acid open-hearth process, due to its limitations, ia decieaaingly in use. THE BASIC OPEN-HEAKTH PROCESS
To make steel by this process, Ume is added to the charge to produce a basic slag, and the hearth is lined with basic material to withstand this basic slag. Iron and scrap steel that contain phosphorus are used. There ate in the United States vast bodies <A non-Bessemer ores yielding a pig-iron too high in phosphorus for the acid open-hearth process, and too low for the basic Bessemer converter, but which the basic open-hearth can remove without difficulty for the production of a suitable st«el.
The method employed for the removal of carbon, silicon, and manganese
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Reactions Basic Open-Hearth Process 335
are the sajne as in the acid open-hearth process, except that in baac practice there is an addition of lime for the formation of a distinctly basic slag which will not attack the basic-lined bottom.
Under the oxidizing action of the flame, and by the addition of some iron ore, the pboBph<mi8 is oxidized to phosphoric acid, and the sulphur is removed as calcium sulphide and manganese takes up a farUier amount as manganese sulphide.
The pbospboruB, carbon, siHcon manganese, and sulphur are eliminated by oxidizing them, the oxygen being obtained principally from the irtHi ore. The reactirais which take place are as follows:
C+FeO=CO+Fe,
Si-|-2Fe0 = SiO2-|-2Fe,
Mn + FeaOa = MnO + 2Feq,
S+2FeO = S03+2Fe.
Of the products of these reactions the CO and SO2 are volatile, and escape as fast as formed, the speedy escaping of the CO causing the boiling of the bath. The phosphoric acid (F20fi), silica (SiOs), and manganese oxide (MnO) separate from the molten iron and unite with any bases present to form the slag, a phosphate and silicate of iron, manganese, lime, magnesia and alumina, the slag floating upon the surface of the bath to be removed by pouring.
CALCULATION OP CHARGE BASIC OPElf-HEARTH CHAROE-SHEBT
FbO+MdO.
C«0+»i«o.
Per Cent
PerC«rt.
Pouod.,
Paimdi,
48
Lunestone
SiOi+PtO(-23,0 per cent FH}+MnO-23,0or 1 SiO,-l FeO+MnO, CBO+MgO-46.0or 1 SiOi-2CaO+MBO Other elementa 8.0
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Wrought Iron And Steel
The Charg«. — Above is given a typical charge to produce a slag ot the compoeition given. The proportions of pig and scrap depend aa the cost, abundance, and relative analyses of the pig said scrap. Both these should be aa low as possible in sulphiu' contents, so that in the pigs, for instance, this should be under 0.05 per cent.
Method of Charging. — The common method is to charge practically all of the limestone, then the pig iron, and lastly the steel scrap. A portion of iron oie is also usually charged with the limestone and the scrap, the heat then has the benefit of the oxidizing action all the time that the metal is in the furnace.
Calculation of the Charge. — Referring to this charge we give the following analysis of its constituents:
Pig Iron.
6t«18ct»p.
Iron Dr..
Trace
Trace
s
The silicon is oxidized to SiOs, the phosphorus to P2O6, the manganese to MnO by the air and the iron ore. The percentage of each element is multiplied byitsfactortoexpressitsamount when oxidized, viz., SiO = 2.1; Si, P205 = 2.3 P; MnO"1.3 Mn. The computations are quite simple. It will be noted, for the slag specified on the charge-sheet, the combined FeO+MnO should be equal to the combined siUca and phosphoric add, the alkaline bases (CaO+MgO) twice that quantity. As figured, the limestone might be increased slightly.
A typical open-hearth charge for a 50-ton furnace is as follows: Molten pig iron from the mixer, 50,000 lb.; steel scrap, 60,000 lb.; limestone, 8000 lb. After melting, the additions in the furnace would be: Iron ore fed in, 1500 lb.; feldspar, 250 lb. (to promote fhiidity). The additions in the ladle are coke, 280 lb. ; ferro-mangancse. 500 lb. ; aluminiun, 1 lb. The ordinary method is to charge all the limestone, then the molten pig and lastly the scrap.
Operation. — ^As shown in Fig. 182, it takes four hours to melt a charge, and six additional hours to complete the manipulation, so that in ten hours the charge is ready to draw, Ihmng the three-to four-hour melting period, the cubon, manganese, and silicon we can see arc reduced. The reactions are ctrntrolled by the melter, who sees that the carbon is eliminated last,
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Reactions, Basic Open-Hearth Process
and if it is oxidizing too fast be must " |Ng up " the charge by the addition of pig iixm to increase the carbon. On the other band, if phoepbonis 18 oxidizing too fast, the oxidation of the carbon cui be hastened by " oreing down " (adding iron ore) to produce the following reaction:
FeaOa+SC-ZFe+SCO.
If carbon is eliminated too soon, much iron becomes oxidized. With the oxidivtion of ^1icon and phosphorus to silica and phosphoric acid, these acids form with liine and iron oxide a basic slag containing 10 to 20 per cent Si02, 6 to 15 per cent PaOs, 45 to 55 per cent CaO, and 10 to 25 per cent Fe. The slag doef! not attack the basic-lined hearth, and retains the phosphorus and the sulphur, but the CaO must be as high as possible for
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Fio. 182. — Chemical Changes in the Basic Open-hearth Furnace.
this, and yet not so high as to render the slag infusible. After meltii^, active oxidation begins, and the bath boils by the escape of gas. Upon the completion of the operation the charge is ready for tapping into a 50-ton ladle, the metal filling the ladle and the l^ht slag overflowing and being thus removed. If the slag remained, phosphorus would be reduced from it, upon addition of the recarburizer, and would again enter the steel.
Recarbmization. — Alike in acid and basic practice, this dignifies the addition of ferro-manganese, containing both manganese and carbon, which restores to the melted charge just enough of these elements to give the desired qualities to the steel. The amount needed to give 0.50 per cent manganese in a heat of 100,000 lb. may be thus calculated for Group I, where there is 0.46 per cent or 460 lb. in the steel. If a ferro of 80 per cent manganese is added in the ladle, when the steel is tapped out, there will be a loss of 25 per cent so that 770 lb. of ferro-manganese should be added. Where more silicon is desired it can be supplied by the use of ferro silicc^.
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Open-Hearth Plant 339
Alloy steels take their appropriate metal. To increase the carbon as a hardener charcoal or coke in paper bags is thrown into the ladle and half of this is lost. Ahiminum in small quantities is also added to quiet the metal and make sound ingots. It takes the oxygen from dissolved iron-oxide and itself rises to unite with the slag as AlzOs.
Charge Composition. — In present practice the chai^ for a basic furnace consists of steel scrap (steel trimmed in the process of manufacture, old steel rails, and steel collected by junk dealers) ; of pig-iron containing leas tbao 1 per cent Si, more than 1 per cent Mn, and up to 2 per cent in P ;
Fio. 184. — Crane and Magaet.
of calcined limestone (quicklime) 8 to 30 per cent of the charge; and of iron ore.
THE OPEN-HEARTH BtHLDINO
Fig. 183 is s sectional elevation. Beginning at the left is the underground hopper, vertical elevator and storage bin for the' coal for the producers that make the gas for the three 200-ton tilting open-hearth furnaces. By a goose-neck pipe the gases pass to the undergroimd system of flues and chambers as already described.
On the elevated track, just within the main building, stands the locomotive which brings in the 65-ton metal ladles. These are picked off their trucks by the 100-ton crane and pom^d into the tilting 20ft-ton open-hearth fiunace as shown. Just beneath the ladle on the elevated working platform is a track on which are brought in the boxes of scrap or pig needed for the charge. The open-hearth metal when finished is poured into a 110-ton bottom tap ladle, and then picked up by the heavy 175-ton traveling crane by which it is tapped into the ingot-molds near the right side of the open-hearth building. One notes the small recarburizing ladle by which the ferro-eilicon, etc. (first melted) and other additions are
made. The supem&tant alag, as it EtccumtilateB, is poured into a sl^ pot, set directly beneath the spout.
The throe open-hearth furnaces each of which has a hearth area of 900 sq. ft. are electrically operated. They are so constructed as to be heated either by producer gas or by fuel oil. Fig. 181 shows the charging machine. The cars carrying the boxes of cold stock, whether scrap iron, steel, or pig iron are set in front of the furnace charging-door. The charging bar of the machine hooks upon a box, lifts it off the transfer car and carries iiito the furnace. The box makes a half-turn which diunpe the load, and the empty box is at once withdrawn. In this figure tbe fixed open-hearth is shown as in Fig. 176: in Fig. 183 is a 50-ton tilting furnace. Beneath the charging floor will be seen a section of the gas chamber leading to the stack.
The casting ladles receive the finished charge or heat which is tapped into ingot molds standing on the pit floor. The pit slag from the furnaces is handled in steam-dumping ladles.
The GaB-producer Building and Stock Yard. — This building (not shown) is parallel to the open-hearth building. It contains nine self-cleaning Hughes pTtxlucers (see Fig. 14). For furnishing basic lining there are also two dolomite kilns and a crusher. The stock yard, where is assembled the iron and steel scrap and the iron ore and limestone, is located between the gas-producer building and the open-hearth building. In tbe stock yards by means of a 10-ton magnet crane. Fig. 184, the scrap is picked up and loaded into the charging boxes. The boxes are then brought into the open-hearth building and placed close to the furnaces ready to be loaded into any one of them by means of the charging machine (see Fig. 181).
The Duplex Process Of Steel Haeiitg
The duplex process, now largely in use in large plants, is generally understood to mean the making of steel from non-Bcsscmer pig iron by a combination of the acid-Bessemer and the basic open-hearth process. The acid-lined converter oxidizes the silicon, the manganese and a certain portion of the carbon of the pig, the amount of the carbon depending upon the practice. The blown meta) is then transferred by ladle to the basic open-hearth furnace where the phosphorus and the remainder of the carbon are removed. The oxidation or burning off of the silicon, manganese and carbon proceeds rapidly in the converter, while in the open-hearth the phosphorus as it oxidizes enters the basic slag which does not attack a basic lining. The duplex process shortens the open-hearth purification by more than five-sixths of the usual period, giving a steel of the same quality as the straight open-hearth process.
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Open-Hearth And Electric Furnace Plant 341
In practice the p^^-iroa is poured into the converter, the blast turned on and the heat blown until in the judgment of the blower the jnctal is of the desired metalloid content. In the case that h^-phosphorus iron is used the blow is stopped when the metal still retains about 1.00 per cent of carbon; while, when treating a low phosphorus pig, the metal is nearly decarbonized. The blown-metal together with 2 to 3 per cent of lime to give a basic slag is now charged to the open-hearth furnace. If the metal has been decarbonized 10 per cent molten p^ iron is added either in the transfer ladle or in the open-hearth furnace. In the furnace a reaction takes place; the phosphorus oxidizes said enters the slag as phosphate of lime while the carbon is removed as carbon dioxide. When the phosphorus is within the specified limits, as determined by a rapid laboratory analysis, the heat (the chai^) is tapped into a ladle where .proper addition for the required manganese and carbon content arc made in the steel ladle. The xisual way is to have three 20-ton converters supplyit^ one sixty-ton open-hearth furnace with metal. If the three 20-ton converters are blown tc^ther and their imited metal assembled in one transfer ladle for removal to the open-hearth, such a plant can keep four or five open-hearth furnaces in continuous operation. This is due to the time needed for the respective purification, the Bessemer taking from fifteen to twenty minutes while the open-hearth will take from ninety to 1 10 minutes.
Fig. 185 gives the general arrangement of the converter and open-hearth departments of a large plant using the duplex process with the addition of an electric furnace building and a forge-press building. It is arranged to provide Bessemer metal for open-hearth refining, and open-hearth metal for electric refining; also directly made Bessemer ingots and open-hearth ingots.
The location of the Bessemer, the open-hearth and the electric furnaces is shown, also the bottom house where the converter bottoms are mEide and dried. The position of the forge press building is also indicated, but the latter is not described.
The Converter and Mixer Building, Fig. 174, shows many details of operation in making steel from Bessemer pig from the receipt of the molten metal from the blastr-fumaces to the production of the ingot.
The molten metal is tapped from the blast-furnaces into a " 65-ton ladle " carried on a truck at the ground level near the converter. It is lifted from the truck by the " 100-ton crane," and poured into the " 1300-ton mixer " where a large body of molten metal is accumulated of the average grade produced by the blast-furnaces. When a charge is needed for any converter, the mixer is tilted and a part of its contents poured into
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Electric Steel Making 343
the " hot^netal transfer car " which travels <m an elevated platform at the height needed to enable it to pour into any ctaiveiter vheu this is turned down into receiving position. To do this, the ladle being in position, a hook beside it tips the ladle. The hook is attached to a steel rope traveling over pulleys and actuated at the converter, so that as the converter turns down the ladle begins to pour. As the converter is turned back blowing begins. Additions of scrap or pig metal are made to the mixer or to the converter during blowing from & concrete scrapping platform above them. The scrap is brought in boxes carried on trucks and these boxes are hoisted from the trucks and delivered to the platform to be added to the charge as needed.
The blown metal is poured from the converters into a bottom-tap ladle which is then brought over the ingot molds standing upon trucks seen cloee to the mde of the building at the left. The ladle is bottom-poured or teemed into the molds. For closer observation in pouring the craneman's cab is carried on a frame secured and braced to the traveling crane. A platform at the level of the top of the molds is for the tapper who does the pouring. Just outside the building and near by is the blowii^ platform where the converter man stands to operate the converter 80 ft. away. The air supply ctxnes from a pressure blower in an adjacent building. When cool, the ingot trucks are taken to the stripper yard, when the molds are stripped or lifted off from the ingots and these are sent to the rolling miU.
Electric Steel-Haeiho
The manufacture of electric steel is becoming well established in the United States, the estimated tonnage for the year 1919 being 1,215,000 short tons. This increase has been due :
(1) To the production of a more uniform quaUty of steel.
(2) To the fact that electric steel can be poured at a much h^er temperature when stiU or dead, ensuring the production of thinner castings.
(3) That the tensile strength and other physical properties show that the steel is stronger and tougher than other steel.
In addition aae has to remember that in the United States in 1919 there were about 85 electric furnaces in the non-ferrous metal trades and about 100 producing ferro-alloys.
The Electric Furnace.— Figs. 186 and 187 are perepective views of an electric furnace and control panel and of the transformer and substation equipment respectively. As shown in the sectional elevatitm, 188, it is a three-electrode tilting furnace, resembling in its action a great arc-light. It uses an alternating current. The current in the substation is transformed from the supply line to a low.-voltage, high-amperage current, thus giving a heavy current for the melting. In the figure a
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clean pit is shown beneath the furnace while, in Fig. 189, the funiace is set so as readily to pour to ladles set on the Boor. The furnace shell is a steel pan having a brick and ganister lining and with a firebrick roof. It
Fio. 186, — Electric-furnace and Fia. 187.— Tranafonner and Sub-
Gontrol-panel. station Equipment.
is charged with steel scrap through a door at the right, and the electrodes are lowered upon the charge heating and melting it. When ready, the slag is poured off through a spout at this side, while the metal is received into a ladle at the left, as shown in Fig. 189.
ing Lining and Bottom Neutral
The Electric-furnace Building. — Centrally in the sectional elevation of Fig. 189 is placed the three-pole electric furnace, having at the ground level at the right a 50-ton steel transfer ladle and on the left a slag pot or car for removal of the slag from the furnace. The whole building is
Elexdtric-Furnace Plant
commanded by & 60-ton traveling crane for chargbg, and removal of the finished metal in a 30-ton ladle to the ingot molds. At the left side is a 5-tion wall crane for stripping. The mechanism beneath the crane ie for electrically tipping it through medium of a sector and spur gearing. The furnace has three carbon poles, as shown in Fig. 188.
Fio. 189. — Electric-furnace Building.
Varieties Op Steel
Basic Open-heartii Steel. — This is a quite pure steel containing less than 0,10 per cent impurities.
Hi^-Grade Steel. — A steel made in the electric furnace. Steel rails are designated as follows :
C.
Si
P. Mn.
Not over 0,20 Not over 0 20
NotoverO 10 0 70 to 1 , 14
Open hearth
Not over 0 04 0 60to0.90
Wbought Iron And Steel
STSBM:ASTraGS HAVE THE FOLLOWWG PROPRRMES
(Eitr«( tram SpBciflwtioM, Anwr. 8«dety lot T<»tiDC UaUtuli)
Per
c.
a
P.
Otdiiury castingB
Tested castiuBB, hard
Tested castingi, medium.. . Tested castinei, soft
None required 38,260 15 20 31,500 18 25 27,000 22 30
STRUCTURAL STEELS FOR BUILDinGS ARE THUS DESIGNATED
atruetuni Steel.
Phospho
Phoephanu, maxinium, opeD hearth
Ultimate tensile Btrength, pounds per square inch. Yield point
Chanctei of fracture
Cold bend without fracture
0 . 10 per cent 0.06 per cent 65,000-65,000 1 Ult. tens. str.
0.06 per cent 48,000-58,000 i Ult. tens. (to.
ffilky 180° to diam. of 1 thickness
Silky 180* flat
Tool Steel Is Of The Following Cohpositioit
Tud«-
sr
Cr.
flul.
TbM.
8U.
O.OIS Trace
Alloy steel may be defined as ordinary properly mehed carixm steel to which have been added ferro compounds of certain rare metab in sufficient though small amoimt to materially modify the qualities of the original carbon steel as shown in the foUowing table.
The alloy steels are divided into two groups, those with one metal alloyed as in (3) and quaternary steels with two metals alloyed as in (7). Thegr have high elastic limit, great strength and toughness. The first two qualities arc enormously increased by heat treatment (quenching and tempering) and the steel still retains great toughness. The moet important of the structural alloy steels are those of nickel, chromium, and vanadium, which by heat-treatraent can be given a tremendous range of strength, varying from 100,000 to 250,000 lb. per square inch.
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Steel And Iron Prices
COHPOSmON OF AILOT STBELS
Lathe tools made from high-speed steel can be run at a speed of 30 ft. per minute, indeed so that the cutting edge shows a just visible red, a speed four times as great as that which ordinary tool steel will stand.
mOH ORE AND PIG-IRON PRICES
Pig-iron, IHttsburg Mail:et la 1920 (Quotations for Carload Lota).—
Standard Bessemer, S29.35; malleable Bessemer, S28.65; bade, $27.15; No. 2 foundry, $28.15; gray forge, $27.15. Standard Bessemer is used for making steel in the Bessemer converter, malleable Bessemer for malleable iron castings, basic for steel suited to the basic open-hearth furnace, foundry for making foundry castings suited to machining, and gray forge for wrtnight-inm. In the Chicago markets both Southern and Northern pig-iron are quoted. The first, from the great iron center at Birmingham, Ala., though cheap, is high in phosphorus. The Northern iron from nearby points is made from Lake Superior ores. Pig iron, cast in sand, is weighed to 2260 lb. for a long ton, the 20 lb. excess being allowance for the sand that sticks to the pigs.
Steel, Pittsburg Maitet. — Bessemer and open-hearth billets are quoted at $38.50. These are ingots 4 in: square by 6 ft. loi^ that are re-heated and rolled into the required merchant-steel bars.
Iron Ores are purchased by guarantee on the part of the shipper that they will crane up to a given standard that generally is based upon the percentage-content in natural condition, thus including the ctmtained
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moistuiB. For Lake Superior ore the prices for 1920 at Lake Erie porte, per l<Hig ton (2240 lb.) were :
Old range Beesemer, 55 per cent iron base S9.45
Old range Bon-BeaBemer, 51.5 per cent iron base 5.70
Mesabi Beeaemer, 55 per cent iron base 6.20
Meeabi non-Bessemer, 51,5 per cent iron base 5.55
On a non-Bessemer ore, as it varies from this, a premiimi of 11.95 cents is paid for each per cent iron over the guarantee, and a penalty or reduction is made of 11,95 cents down to 50 per cent iron, and 17.95 cents down to 49 per cent iron, and a double penalty down to 48 per cent iron. Below 48 per cent, the penalty becomes 27 cents per unit. On Bessemer ore, provision is made for a premixun only in case the oi« exceeds the guaranteed 65 per cent iron.
The Old Range ores come from the iron ranges on the south side of Lake Superior, and command a higher price because of the better mechanical condition. The Mesabi ores are soft, friable, and carry much fine, which makes flue-dust. When smelting such ore, in ratio of 85 per cent soft to 15 per cent hard ore, as much as 6 per cent flue-dust or dirt is made.
Steel WoAs, Scrap, 1920. — Heavy melting steel per gross ton deli^'ered at works $19.00 to S21.00.
Cost of Pioductioii of ng-iron in the electric furnace in 1915 was $26.21 per long ton and of steel from steel scrap (29.90 per long ton based on a cost for common labor of S2.50 per eight-hour shift.
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Part V Copper
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Chapter Xxvii
Coppes. Ores And Their Treaihert
CHARACTEKIsnCS OF COPPER ORES
We aie to think of copper ores as mineral aggregates, carrying frequently 10 to 15 per cent copper or leas, with associated minerals and «a earthy gangue. Treating the (»e is a problem not only of obtaining the copper, but of separating and diminating the gangue and associated minerals. Though there are many kinds of copper ores, those of conmiereial importance are few in number. We may divide them into three classes: (1) the sulphides; (2) the oxides, including the carb<niates and silicates; (3) ores containing native copper.
Sn^hides. — Chalcopyrite, CuFeSa, when pure contains 34.fi per cent copper. This is by far the most widely distributed and most abundant of the ores of copper, and furaishes the world's principal supply of the meta). It is frequently accompanied by iron pyrite, and has sihcious gangue even when tlie sulphide is massve. In consequence, the ore often carries no more than 3 to 4 per cent copper, but is particularly suited to pyrite smelting, that is to smelting with little fuel. Silver and gold are found in the ore in small quantity. The deposits at Mt. Lyell, Tasmania, that have been so successfully worked by pyritic smelting, are cluefiy of maanve ircn pyrite, containing chalcopyrite, and canying 4.S to 6 per cent copper with 0.15 oe. gold and 3 02. silver per ton.
Chalcocite (copper glance), CusS, is cinnputed to contain 79.7 per cent copper, but it is seldom pure even in the crystalline form, the copper having been replaced by iron and other metals. The impure mineral shows the characteristics of the pure mineral when carrying as httle as 56 per cent copper. The pure crysti^ resemble the artificial product " white metal," a high-grade copper matte produced in the furnace.
Bomite (peacock ore), CusFeSs, when pure contains 55.6 per cent copper. It is found associated with chalcopyrite and chalcocite in proportions varying from 42 to 70 per cent copper, without losing the characteristic varied colors.
Enargite (SCuaS-AsaSe), 48.3 per cent copper, an arsenide, occurs in Butte, Mont., ons.
Tetrabedrite (gray copper fablers), _ (Cu3S,FeS,ZnS,Ag2S,PbS), 351
352 Copper Ores And Their Treatment
(SbjSstABaSs), may be computed as containing 30.4 per cent copper, but it varies greatly in the copper and silver content. It has already been mentioned as a silver ore. Because of the contained arsenic and antimony it is unfavorable as a copper ore, and it is only because of the richness in silver that it ia treated.
Oxides, Carbonates, and Silicates. — These ores are the result of the decompoeitioii (^ the copper sulphides, by air and water. We find them in the upper zones of mineral deposits accompanied by iron oxide, which also is the result of the decomposition of iron sulphide. As we sink on the vein we find the oxidized ore of the upper levels giving place in depth to the imaltered sulphides.
Cuprite (red copperKixide), CuaO, 88.8 per cent copper, is a product of
decompoeation. It often permeates U-rgfi masses of iron ore. Large lumps
^ of the ore are sometimes found, the center of which contains unaJtered
metal. These evidently are the result of the oxidation of a mass of ua^ve
copper,
Melaconite (black oxide of copper), CuO, contains when pure 79.8 per cent copper. The ore, with the copper in part replaced by oxides of iron and manganese, ia sometimes found in masses large enough to pay for extraction, and containing 20 to 50 per cent copper. The so-called black oxide of the Blue Ridge r^on, on the border of Teiuxessee, North Carolina, and Virginia, seems to be an intimate mixture of copper glance, black copper oxide, copper carbonate, and native copper with iron oxide and sulphide. The ore can be readily roasted in lump form.
Malachite, CuCOs ■ Cu(OH)a, 57.3 per cent copper, occurs widely distributed, ordinarily in non-paying quantities as a decomposition product in surface deposits, but sometimes sufficiently rich to work. It ia found mixed with limestone, dolomite, oxides of iron, manganese, and silica. It is difiicult to juc^ the copper content of the ore from the appearance, but the green color makes its presence readily recof^iizable.
Azurite, 2CuC08 ■ Cu(0H)2, is comjmted to contain 55.2 per cent copper. The ore is blue, as the name indicates, and the appearance is strildng. It occurs in the same way as malachite, and often is associated with malar chite, but it is lees abundant, and often is only a coloring on other oxides.
Chrysocolla. a hydrated silicate of copper, containii^ when pure 40 per cent copper, is a decomposition product of copper sulphide, and is often accwnpanied by malachite.
Native Copper. — Native copper is found extensively in the copper region of Lake Superior, Elsewhere it occurs sparingly and is not commercially important, though it often accompanies the oxidized ores. In the Lake Superior region it is found in wide lodes disseminated through the lode-matter 0.65 per cent to 4 per cent of the whole, and even when the lowest grade mentioned, by concentrating can be recovered at a profit. The COQ-
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Treatment Of Copper Ores 353
centrate or " mmeral," as it is locally named, is produced in different grades, ranging from 30 to 94 per cent copper. Much of the native copper is pure; in other instances it carries a little arsenic.
Properties of Copper. — Its melting point has been established at 1083° C, and its specific graWty at S.89, while its. latent heat of fusion is 43.3 calories, and its specific heat at 170° C. is 0.09244 closely one-tenth that of water. Traces of oxygen are purposely left, even in the highest grades ci copper, ranee otherwise it would be impossible to cast a sound it^ot, as the copper in the refining process readily absorbs gases that are racpelled during solidification. Even the best of copper casting are somewhat porous, and thus low in electrical conductivity. Castings of a conductivity of 97 Mathiessen's standard have been made by the addition of smaJl amounts of boron in the ladle just before casting, thus deoxidiang the metal so that its mechanical properties are excellent, and it can be used for making even intricate castings.
Copper is mechanically improved by hot working: When it is heated to a bright red and quenched, maximum ductihty is attained, while cold working increases the tensile strength, but lowers the ductility.
Solid copper can absorb arsenic up to a maximum of 4 per cent. In small quantities it increases its maximum stress without affecting the ductility. It has been found that the deoxidation of arsenical copper by addition of ferro-silicon greatly improves its qualities.
The physical properties of copper fall into two claseee, viz., electric and mechanical, and the treatment best suited to attEun the one is undesirable for the other. Pure, soft dense metal has the highest conductivity, but it is weaker for use on transmission lines.
THE EZTRACnOn OF COPPER FROM ITS OSES
Copper may be extracted from the ore by dry or by wet methods. By the dry method the ore is smelted, the process being one of igneous fusion. By the wet or hydro-metallurgical methods the copper is leached ' from the ore. The striking point of difference between the two methods is that, in the first, we melt the entire ore, effe<!ting then a separation of the copper frton the worthless part, while in the wet method we act upon the copper alone, leaving the greater part of the ore in the original condition.
The Diy or Igneous or PyrometaUui^ical Methods. — Probably more than 90 per cent of the world's production of copper is by smelting. The methods of smelting vary with the nature of the ore. We may divide them into the following:
(1) The Smeltbig of Oxidized Ores that may contain a little copper, or of concentrates of native copper in blast-furnaces and in reverberatory furnaces is done for the production of a crude copper called blister copper.
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364 Copper Ores And Their Treatment
When performed in the blast-furnace the process reaembles the smelting of ircai ore to produce pig iron. For fine ores or concentrate the reverberatory furnace is preferred, Butce in the blast-fumaoe much flue dust is produced.
(2) The smelting of ores containing sulphides of copper and iron in the blast-furnace <» the reverberatory is to yield a copper-and-4ron sulphide, called " matte." The afnount of matte is dependent on the amount c^ sulphur per cent, so that if part of the sulphur is expelled l^ roasting, or by being burned off or volatilized in the blast-furnace, then lees matte is formed. As much as 70 to 80 per cent of the sulphur may be got rid of in this way. The copper is concentrated into a small amount of matte as c<Mnpaied with the original bulk of the ore. This matte has to be further treated to obtain Meter copper.
Ordinary Matte Smelting. — ^When roasted ores are smelted, this may be called ordinary matte smelting: About 10 per cent of the charge is coke, added to cause its melting. If the ore is imperfectly roasted or is smelted raw, then less coke may be used, since the sulphur in the ore bums, producing heat; also, this burning c^ the sulphur is in itself a kind erf roasting. Thus, again, the lunount of matte produced is materially reduced.
Pyrite Smeltisg. — The smelting of raw ore in the blast-fuinace is called pyrite smelting because there is much iron or copper pyrite in the ore.
Collectors. — The blister copper produced in the oxidisii^ smelting is to take up or collect within itself any gcAd or silver present in the ore; the same is true of the matte; so that we say both Ulster copper and copper matte are collectors of gold and silver.
In reverberatoiy smelting the elimination of sulphur is less, say 25 per cent of the sulphur in the charge; so this way of smelting is not suited to the treatment of raw ore. Much material for Knelting has been concentrated and so is fine. It is in good condition for cheap roasting in one of the mechanical roasters already described.
Ffaie Concentrates. — ^This fine product, if treated in a blast-furnace, would produce much flue dust; so for such material, the reverberatory is preferred, due to its large capacity, its quiet conditaou erf smelting, and to the cheaper fuel needed.
The Wet or Hydrometallurgical Methods. — In these the copper is obtained frcun the crushed ore in water solution, either with or without the aid of other solutions. The ore may have first to be roasted. From this water solution the copper is precipitated, melted and refined.
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CHAPTER XXVIII COPPER BLAST-FUItNACE SMELllNG OF OXIDIZED ORES
This resembles the smelting of iron ores in that metal is obtained in metallic fonn in one operation. The ore, which does not contain sulphide, is charged into a blast-fumaGe and smelted with coke for fuel. The products are slag and blister^opper, the latter being metallic copper containing impurities that have been taken up in smelting, much as cu'bcm and siticon are absorbed in iron smelting.
BlftBt-funuce Plant for Oztdized Ores.— F^. 190 is a plan and P^. 191 an elevatdou of a blaat-fumace building suited to the smelting of oxidized copper ores. It has two Qoors or levels, the upper, called the charge-floor, and the lower, the slag-Soor. The ground at the right drops away and furnishes a place for a dump. The ores and fiuxes are stored in bins on the groimd at the charge-floor level, and are brought in weighed charges to the furnace door, the siU of which is Bush with the charge^oor. This door is shown also in Fig. 65. The slag and blister-copper are withdrawn near the bottom. Behind the Furnace is seen the pipe or blast-main by which air is conducted to the wind-box at the tuyeres. The furnace stack extends above the roof and at the side branches to a down-take leading to a dust-flue. Here much of the dust, as in the dust-catcher of the ircm blast-furnace is removed. This dust-chamber terminates, as shown in the front of the plan view, at a stack which takes away the residual gases at a bi^ level. On the plan we also see the boiler and engine which drive the furnace blower at a pressure of 12 to 15 oe. per square inch, equal to 24 to 30 in. of a mercury coliunn. Waste slag, or sweepii^, carrying copper, are returned to the feed-floor by a platform elevator shown in the comer of th« furnace-room.
In Fig. 65 is a view of the cupola blastfurnace used for the production of copper from oxidized copper ores.
The crucible contains the molten contents of the furnace, the copper below and the lighter slag floating upon it. There are two tap-holes and two spouts; the lower, close to the bottom, is to remove the molten copper; the upper, a few inches higher, is to withdraw the slag. From time to time, as slag or copper accumulates, it is withdrawn by piercing a hole through the clay-stopping of the tap-holes by means of a pointed steel tapping-bar. The flow is arrested by thrusting into the opening a plug of clay stuck m 365
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Copper Blast-Furnace Smelting Of Oxidized Ores
Tin. 190.— Sectional View of Small Smelling Pl&ut.
,
■m
L.I
Pj|
lU
Fio. 191.— Plan of Siaail Smelting Plant.
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Copper-Smelting Plant 367
the end of a button-headed stopper-rod or dolly. The slag is received into a fore-hearth movinted on wheels through which flows the escaping furnace slag. This s\»% carries with it drops of copper not settled out in the furnace. The molten slag quite fills the fore-hearth, crusting over, but maintaining a cavity, where the drops of copper settle out. The slag over6ows at the spout at the opposite end into a slag-pot, Fig. 200, set to receive it. When the tap-hole is stopped by a plug of plastic clay, the slag flow ceases, and an empty slag pot replaces the full one.
The copper is received ma" bullion mold," which, after filling, stands until the copper has solidified. The ingot is then dumped out and the mold again used. The furnace shown is a round one, 36 in. diameter inside, thu8 having a bosh or enlargement of 6 in. on the side.
In operation, the furnace is kept full to the feed-door with alternate layers of fuel and charge. The blast rises through the column of materials (a distance of 7 ft.) and passes off through the down-take, which has sufficient draft to take away the gas and smoke and also the air that enters the feed-opening or door. The blast enters under pressure causing an intense combustion of the coke and the fusion of the chat^. The copper reduced by the glowing coke, collects in drops and finds its way to the bottom of the crucible, while the gangue of the ore, fluxed by the addition of iron and limestone, forms a fusible slag.
The smelting of the oxidized copper ores, as above described, was formerly used in the southwestern United States so long as the oxidised ores lasted; it has been replaced by the more eflicient and cheaper methods ci matte-smelting either in blast-furnaces or in reverberatories.
The notable exceptions are those of the copper country of Northern Michigan and the large-scale work of the Union Mini^re du Haut Katanga, Southern Congo.
Both tne cupola furnace (Fig. 65) and the rectangular furnace resembling Fig. 196 are xised. The smeltii^ is conducted along the Unes described for oxidized copper ore for the production of an impure blister copper. Reverberatory slags to which has been added native copper and briquetted material are thus smelted. The charge consists of slag 2000 lb., limestone 600 lb., and anthracite coal (steamer size or about 2} in. diameter) 400 lb. It is the practice at one works to add much small mass or native copper to the chaise, the idea being that, as the native copper in it melts and sinks to the crucible of the furnace in the form of drops, it carries down reduced copper with it. Elsewhere the fine concentrate has be^ thoroughly incorporated with a quicklime paste, briquetted in a briquetting
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358 Copper Blast-Furnace Smelting Of Oxidized Ores
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SMELTING AT UNION MINli:RE DU HAUT KATANGA 35ft
press and the briquettes treated in cylinders imder steam pleasure for twenty-fo\ir hours. This renders them quite hard, and in an acceptable form for smelting.
The products of the furnace are 8]&g of less than 1 per cent copper and " cupola blocks," an impure blister copper, which is sent to a rererberatory furnace forreSning.
It is to be noticed that the low percentage of copper in the slag is due to the use of the anthracite coal. This has an intenser reducing action than the coke. However, it slows down the furnace and the two fuels are to be used in judicious proportions, beoause, as the anthracite is cut down and the coke increased, the furnace will run the faster. This use of coal suggests itself in cases where oxidized ores have elsewhere to be smelted in a blast-furnace.
The cupola Mocks, referred to above, which constitute the product of tbe Uast^umace anelting of the slag frcon the reverberatory re&ningfumace, are melted and refined to produce a low-^rade copper called " casting-copper." The cupola-blocks are impure and contain so much arsenic that it is practically impossible to remove it all. Other impurities (iron and sulphur) are eliminated.
SMELTING TO BLACK COPPER BT THE TUnON HUnBRB DU HAUT EATAHOA
The company smelts oxidized copper ore from the great superficial deposits at Lubumbashi, near Elisabeth ville, Belg^ Congo, producing black or blister copper as already described under the heading " Blast-furnace smelting of oxidized ores." There are six furnaces, Fig. 192, each 44 in. wide and 20 ft. long, the largest thus far made for the treatment of oxidized copper ores. They have two tiers of jackets, the lower side jackets being 10 ft. hi(^ and 2 ft. wide. The upper ones of the some width are 7 ft. 4 in. h^h. From the tuyeres to the feed'Boor is 18 ft., and the ude-bosh is 14 in., making the furnace shaft above the lower jackets 72 in. wide, for good reduction. At the left-hand ^d is a trapped slag-spout for the ccAciuuous flow of the slag ; at the other end is a slag-spout to be used in case the furnace stops and the slag has to be tapped off. At one side, and Dealer the right-hand end, is the bullion spout and tap-bole for removing the bullion (black copper) from the bottom of the crucible. To the top of tbe crucible from the slag-floor it is 7 ft. 2 in., so there is suflicient room for the fore-hearth beneath the trapped slag-spout. The bustle-pipe, 21 in. by 42 in., in cross-section ia rentangular in order to save building space. The furnace is surmounted by a closed top (not shown) as m Fig. 196, with feed-doors the full length of the aide.
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Chapter Xxix Blast-Fdiutace Smelting Of Sulphide Ores
Matte Smelting
If we smelt raw ^phide ores of copper and iron in the blaat-fumace just described, iising, say 10 per cent of fuel, with added fluxes to form a fusible slag, we form an artificial sulphide or matte of 23 to 25 per cent sulphur. The copper that was in the ore and tUe iron from the charge enter the matte, but the qxiantity of matte so formed is but little less than that of ore originally put into the furnace. If, however, we first roast the ore, the quantity of sulphur present, and consequently the amount of matte made, is less, and the ratio of the ore to the matte may be five or ten to one. In the matte will be the copper ae a sulphide, and, in forming, the matte will take up the silver and gold of the ore. By this operation we cc^ect the precious metals in a product one-fifth to one-tenth the original ore, the matte being termed a " collector." It then can be further treated to convert it into metallic copper carrying the precious metab. By the the process of electrolytic refining, the gold and silver are eventually separated from the copper? A charge suited to matte-emelting methods would therefore consist of roasted ore retaining 7 per cent sulphur tt^ether with oxidized copper and copper-free ores containing gold and cdlver added to recover the precioua-metal content. It would also carry fluxes to make a fusible slag and to supply iron (if needed) for the matte.
The products of the furnace are sl^ and matte. The former is the result of the imion of the silica in the chaige with the various bases, chiefly iron oxide and lime. The latter is the complex artificial sulphide produced by the sulphiu in the charge combining with copper and iron. The affinity of sulphur for copper is greater than for iron, and it takes the former first; then if it needs iron it takes that also, until a compound of both has been formed that contains approximately 25 per cent sulphur. Any further iron present enters the slag as ferrous oxide, and as the ratio of the iron thus available to the other bases varies so will the slag vary in composition; but the principal requirement is that the quantity of silica be enough to form a fusible slag. Slags of 25 to 40 per cent siHca are common in copper-matting practice, and the 40 per cent limit is sometimes exceeded when, for economy in smelting, it is desired to use as little flux as possible.
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Copper Matte Smelting
In the matte-Eunelting operation the object ia to collect the metslfl, copper, gold, and eilver, into a small amount of that complex artificial iron-copper sulphide called matte. Now, in the pyritic Bmelting of copper ores, using an excess of air and little fuel, much of the sulphur (say 75 to 80 per cent) is dissipated or volatihzed so that a small amount only is left to form matte. The same result is attained by first roasting the ore, only in such case much more coke must be xiaed, as compared with that needed in pyritic smelting. Blast-furnace smelting is suitable to ore in liunp form, whether unroasted, or as the product of heap-roasting.
Fia. 193. — Meeaiter Bedding System
Ores for Matte-smelting. — Ores suitable for matte-smelting are the oxidized ones containing some sulphur and ores that have been roasted; to which may be added silicious and oxidized ores containmg gold and silver. Ores that require to be first roasted are better roasted in lump form in heaps or stalls or sintered, for the reason that the blast-furnace, because of its strong air currents, is not suited to smelting fine ore. The amount of matte made (matte fall) depends upon the quantity of sulphur in the charge, and to get sufficient concentration (little matte from much ore) the sulphur is kept low. To the ore above described is added limestone and iron ore as flux, and a quantity of fuel equal to 10 to 15 per cent of the charge.
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THE HESSITBR SYSTEM OF BEDDniG
Instead of Btonng ore in pockets, or bedding upon the ground as above described, the Mesdter system of bedding and reclaiming ore is coming into increasing use for lar^ plants. As shown in Fig. 193, it constitutes part of a complete belt-conveying system. The ore, coarse-crushed at the sampling and crushing mill, is there separated by trommels into fine and coarse ore, the fines going to pockets for roasting.
There are three bedding floors dde by side upoa which the coarse ore
Flo. 194 —Robins- Messilar R«clainuiig Machine.
is deposited, each bed whmi completed lieing of triangular shape in croBssection, 375 ft. long, capable of holding 10,000 tons. By an incline conveying belt the ore is delivered upon the cross-conveying belt. No, 1, at the left-band end to the spreading belts Nos. 2, 3, and 4, by means of a tripper, each of these three extending from end to end of the bed. A tripper on each spreading belt travels back and forth over the extent of the bed at the rate of 400 ft. a minute, dropping its load in thin layers upon the bedding floor beneath. In this way lot after lot is distributed imtil the bed is completed.
To take up or reclaim the ore, a reclaiming machine Fig. 194 is used. It is a traveling frame, R, spanning the width of a bed, and a trench containing a conveyii^ belt that takes away the ore delivered to it by the
Co Otitic
Messiter System Of Ore Bedding 363
leclaiming machine, as shown in Section A. Under the forward edge of the bridge is a scrapeiMK)nveyor, operating in a steel trough that hae a flat bottom and a vertical back plate. The flights of the conveyor, sweeping along this trough, carry the ore to the end of the machine to drop it ui>on the trench conveyor-belt. A triangular harrow covering the croaa-
Fia. 195. — Copper-mattiDfc Blast Furnace.
section of the bed and set at a proper inclination has a slow but powerful action back and forth suflicient to dislodge the material which then rolls down within reach of the flights below. Both harrow and flight conveyor are motor-driven. Another motor advances the bridge into the bed at the desired rate or may move it backward, when it is transferred
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364 Blast-Furnace Smelting Of Sulphide Ores
to another bed. This is effected by a long transfer car (see Fig. 303) moving in a tieucb at one end.
THB COFPER-UATTinG BLAST-FDSKACX
For producing matte from copper-bearing orea, whether these are to be smelted .after roasting, or treated raw, by the method of pyritic
Fia. 196.— View of Copper blast-Furnace.
smelting, we \ise the furnace, Fig. 190, aJready described, or one of the rectangular type, Figs. 195 and I9C.
Fig. 195, at the left, represents a transverse sectional elevation of a furnace of 42 by 120 in. interior hearth-diniensons, having eighteen tuyeres, nine at each side, and a capacity of 150 tons of chsugc daily- Fig. 196 is a perspective view of a similar but larger furnace, differing from Fig. 195 in having a trapped slag-spout, more fully shown in Fig. 196.
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Copper Blast-Furnace 365
TTie aole-plate of the fxirnace rests on jack-acrewa, and can be lowered and act aside when it is deaired to make repair or to clean out the furnace. It is protected from the action of the molten matte and slag by a 9-in. lining of firebrick. In Fig. 195 are seen crucible plates which rest upon the solo-plate. These are lined with 18-in. of brick. The hollow jackets filled with water, shown in Fig. 196, extend down to the sole-plate and the water-cooling is sufficient protection from the action of the molten materials. The sole-plate within the furnace, however, is covered by the brick lining. The jackets, shown separately in Fig. 197, are at least 9 ft. high, and in the furnace repreaented.there are two of them on each side, and one at each end. At one end the jacket is shorter, and the space below is filled with a water-cooled tap-jacket through which the slag ia withdrawn. In Fig. 195 the longitudinal view shows the arrangement of a furnace with three
Fig. l!)7. — Water-japkelfl for Copper- Matting Blaal-furn&oe,
jackets at each side, and two jackets at each end. The small jackets are easily handled and replaced. In Fig. 197 the inlets for water are at half the height of the bosh, and the water outlets are at highest point to keep them full of water. They are tied or clamped together with heavy angles, but in the other figures with I-beams.
A water-cooled trapped spout ia used in connection with the furnace as indicated in Fig. 196. Through it flow the slag and matte. Before the slf^ can overflow it must fill the spout and cover the outlet or tap-hole through the jacket. It thus prevents the escape of the blast, and flows in a regular stream as fast as it forms within the furnace. The furnace shown in Fig, 195 is arranged differently. There is a spout and a water-cooled tap-jacket at the end of the furnace through which the slag b removed, while the matte, as it accumulates, is removed by a spout and a side tap-hole at the level of the crucible-bottom. In this case, the separation
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366 Blast-Furnace Smelting Of Sulphide Ores
between the slag and matte is affected within the furnace; in the former case, where the trapped spout is used (ednce matte and slag issue together) they are separated outside the furnace in the foie-hearth or pettier. Fig. 199 or 195.
The transverse view, Fig. 201, shows tim Bide-jacketB. These ba've brackets or knees riveted to them and rest cm I-beams that aro secured to the colunma. Thus, when the sole-plate is removed, the jackets lemain in place. The distance between the side-jackets is 42 in. at the tuyeres, and 66 in. at the top. The bosh, or enlargement, is thus 19. in. on the aide. Above the jackets are the cast-iron distributing plates, forming the sills of the feed-doors. The feed-doois in the opposite long sides of ihs furnace make it aoceesible from end to end, not only for feeding and trinuning
Fig. 109.— Portable Forehearth or Settler.
the chai^, but for cutting with chisel-bars the accretion or scaffolding that may form on the interior surface of the jackets.
The portion of the furnace above the feed-floor level, called the stack or top, is of brick supported by a deck-plate or mantel-plate of I-beanu resting on the cast-iron columns that extend down into the foundation. The upper portion of the stack is a hood of sheet-steel tenninating in a pnpe that extends through the roof of the furnace building. Sometimes a branch pipe leads from the hood to a dust^chamber where the dust w collected.
The bustle-pipe, by which the blast at a pressure of } to 2 lb. is brought to the furnace, extends around three sides and connects by the sheetmetal pipes to the tuyeres. The tuyere ia 6 in. diameter and has a 6-in. screw-cap into which is inserted a nipple with a cap having
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Accessories Of The Blast-Furnace 867
a mica-covered peep-hole, through which the condition of the furnace can be observed. At the branch above the tuyere is shown a slittevalve. In Fig. 196, just below the busUe-pipe, is a waste launder to receive the ovbtAow from the jacket, and below it is a 3-in. water-supply iMpe branching to each jacket, and to the water-cooled trapped Bpaut at the front.
Accessories Of The Blast-Fdkhacb
The fore-hearth, made of cast-iron plates, 4 by 6 ft. inside dimensions, is lined with a layer of brick, and is mounted on wheels bo that it can be quickly set aside and a new one put in the place when needed. The slag and matte flow into it at one end and keep it full of molten slag. At the other end the slag flows out. The matte settles on the way and collects in the bottom of the fore-hearth, and, when accumulated, is tapped at the tap-hole and spout seen at the side. Meanwhile the slag, flowing from
Fia. 200. — Two-wheeled Slag-pot.
the fore-hearth, is caught in alag-pots. Fig. 200, and taken to the edge of the dump and poured. The slag cools on the surface of the fore-hearth and forms a crust from beneath which the molten slag flows. Crust forms also at the sides and bottom, and becomes gradually thicker; and after several days becomes so thick that the molten part of the interior is too small to permit of a good separation of the matte from the slag. When this results, the fote-hearth is pried back on the wheels, and replaced by another. In the furnace, Fig. 196, at the middle side-jacket, another tap-hole furnished with a spout is seen. This is generally kept closed, but is opened when it is desired to empty the hearth of the matte and slag.
The Slag Pot, as shown in Fig. 200, is used for small furnaces both in making bhster copper and for matting. It is hand-drawn and emptied at the edge of the dump.
Ladle Cars. — These, are used for slag and matte in the operation of a lai^ furnace and, as seen m 208, may be in a train drawn by an industrial locomotive.
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368 Blast-Furnace Smelting Of Sulphide Ores
Fig. 201 is the type of fore-hearth used for large furnaces. It laste indefinitely. It is lined on sides and bottom with basic biick to resist the corrosive action of the sl^. The slag flows into it at the back from the trapped fumsce spout and escapes in a steady stream at the slag spout in front, to be taken away in a slag pot to the dump, see Fig. 210. At the side is seen the matte-tap whence from time to time the matte is tapped out into a lai^ ladle and taken to the converter.
Pra. 201. — Stationary Fore-hearth.
Flo. 203.— Positive-presaure Blower,
Blowers. — For furnishing the blast to the furnace the positive blaat rotary blower is used, as shown in Fig. 203, and in section at 204. It will be seen that, by the rotation of the two impellers meshing into one another so that the air cannot escape backward between them, the air must be delivered to the furnace in a positive manner and not, as in a fan blower, be able to escape backward when the pressure rises sufficiently. The air delivered is reckoned at the displacement per revolution. If the pressure is greatly increased there is a backward leakage of air,
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Reactions In The Copper Blast-Furnace 369
caUed the slip. Two boxed-in cut gears outside the blower keep the impellers exactly in mesh. Blowers are often direct-driven by electric motor; but the one described above is belt-driven.
Blast-Furnace Conditions
The diagram, Fig. 205, shows the reactions and conditions within the shaft when a furnace is smelting roasted ore with a full amount of coke. It will be understood that the furnace is full to the top with solid charge and the diagram shows the course of the descendii^ charge and of the rising
Ittte CX>, 'itmi' Neutral atmospbate
in Pyritic Smelting.
Large Copper-Hatting Blast-Furnaces
The tendency of late years has been to increase the size of copper-matting blast-furnaces. Increase in width would require higher blast pressure to drive the air to the centerof the furnace; hence increased capacity had to be gained by increasir^ the length of the furnace. At the same time, to supply more air, the blast-prcsRure has been increased in some cases to 40 oz. or 2^ lb. per sq. in. A furnace 56 by 180 in. under these conditions smelts 400 tons of ore daily. For the matte to properiy settle from the slag, with so large a flow, a cylindrical fore-hearth. Fig. 201, has been used, 16 ft. diameter by 5 ft. deep exterior dimensions, lined with hard firebrick; for a low^rade matte, basic brick is sometimes used. In the crusted-over pool of molten slag, the separation is effected. The molten contents of the furnace flow into it at one side and the slag flows out at the opposite side. From time to time the fore-hearth is tapped at
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Blast-Furnace Smelting Of Sulphide Ores
Pia. 200. — LoDgitudinal Se<^ion of Copper-matting Blost-fumacc.
—Transverse Section of Copper-matting Fiirriibfc.,-.QT,,
Operation Of The Copper Blast-Furnace 371
the lower tap-hole, and 5 to 10 tooa of matte are dmwn into a ladle, for further treatntent at the converter. The lengthening of the furnace has been carried ao far that, at the Washoe plant, Anaconda, Mont., a furnace 51 ft. long, having 1600 tons daily capacity, has been for some time in operation, and recently one 87 ft. long and 3000 tons daily capacity has been built and operated. The firat furnace has two fore-hearths, each 16 ft. diameter and the second one three of that size.
Figs. 206 and 207 are elevations cf a type of copper blaetrfumace of the Dew plant of the Granl^ Cons. Co. at Anyox, B. C. It is 54 in. wide at the tuyere level and 30 ft. long. The crucible is supported upon a water-cooled base-plate, and its slopit^ bottom slants both ways to the trapped spout at the side. The jackets are in two tiers. The space between them widens to 6 ft. above and at the throat this is contracted to 4 ft. 6 in. At each side are feed-boxes and at the right is seen a chai^e- car dumping its load into one of them. When ready the charge is pushed into the furnace by a- cylinderH>perated plunger that forms the back of the box. It falls to the top erf the charge several feet below. By varying the speed of the plui^r, the cascade of ore can be made to fall nearer or farther from the opposite side, as experience suggests.
Regular Operatioh Of The Copper Blast-Puiuiacb
Stajttng Oie Furnace. — Since a blast-furnace is water-jacketed the operation of warming it is a simple one. The firebrick lining of the cru> cible is dried and warmed by several hours' heating with a wood fire. The end and side tap-jackets are removed to permit the ur to enter to the fuel. When the hearth is hot the wood ashes are scraped out and a fresh fire of wood, filling the crucible a foot deep, is started, wood of imiform sized pieces for uniform burning being selected. Upon the wood is placed charcoal, and upon the charcoal coke, until the surface is IJ to 2 ft. above the tuyeres. The fire is increased uniformly and regulated l^ checking the draft at tbe front and admitting air at the rear as required. When the coke is thoroughly ignited, the furnace is ready for charging. The brick-lined fore-hearth, Fig. 201, is warmed while warming the furnace. The wood is placed carefully against its walls, leaving the center clear for the air to reach the fuel, so that the burning may proceed actively. As the wood bums, charcoal and ashes accumulate, and are shoveled out, since otherwise they form a layer through which the heat does not penetrate.
Suppose the chai^ of ore and flux to be 2000 lb., and that we intend, as in regular wori, to use with it 12 per cent coke or 240 lb. per charge. We put in a layer of 240 lb. coke, then one of 500 lb. slag. This is followed by a half dozen charges each of 240 lb. coke alternated with 1000 lb. slag. Next we put in a half dozen charges of 240 lb. fuel and 2000 lb. slag, so
372 Blast-Furnace Smelting Of Sulphide Ores
that the slag when melted ^all entirely fill the fore-tiearth. We now begin feeding the regularly calculated charges sod required fuel. At this time the blast is admitted, gently at first, and increasing during a half hour, after which the furnace should be in full blast. Extra men should now assist in charging to rapidly fill the furnace. At the slag-floor, before the blast is turned on, the tap-jackets and the tapped spout are put in place, and all openings cloeed with a clay phi^ing mixture. This may be obtained from a neighboring bank if of suitable quality, or may be made frtsa coarsely ground fire brick mixed with clay.
As the smelting proceeds, by looking into the tuyeres, we see that the slag is rising to their level. We then open the tap-hole and permit tbe slag to flow into and quickly fill the fore-hearth. The excess steadily overflows to the slag-pots set to catch it, or it may be granulated and removed by water.
When feeding the furnace, cant b taken to distribute the charge evenly, not feeding coarse ore in one place and fine in another. Unless we exercise care in this regard we have irregular operation, blast and flame coming up in one place and the charge looking dead in another. When this begins to occur we load the active places with charge, feed lightly, and »ise coaiaa material where there is little action.
We may adopt either the intemiittent or the continuous method of removing the slag from the furnace. In the intemiittent method, as arranged in Fig. 195, the slag is tapped from time to time as it accumulates as already described, taking care that it does not gather in such quantity as to rise to and run into the tuyeres, or to " slag " them, as it is called. By the continuous method, the slag and matte flow continuously from the furnace through the trapped or open spout, see Fig. 196. The molten products enter a fore-hearth and the separation of slag from matte is there made.
Referring to the view of a smelting plant, Figs. 190 and 191, the furnace being filled to the feed-doors, we have a 7-ft. smelting column (distance from tuyeres to feed-door). As the charge smelts and the molten msteriala are withdrawn, the surface gradually sinks, making room for further additions. The coke is first added in a layer over the surface, and upon it is spread the weighed charge. The air, under pressure from the blowers, driven into the furnace at a prc;«ure not less than | lb. or 12 oz. per square inch, hums the descending coke mostly at the tuyeres, The resulting gas with the sulphur dioxide froni the burning sulphur in the charge appears as a whitish smoke mingled with dust mechanically carried. This pases from the fumace-top directly into the air, or to a du-st-chamber and thence to the stack. The sulphur remaining unites with the copper and a part of the iron and forms a matte or copper-iron sulphide. The matte, in forming comes into contact with the gold and silver contained in the ore of tbe
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Disposal Of Slag And Matte 373
charge and abeorbs them. The coke reduces the iron not needed for the formation of the matte to ferrous form, and the ferrous iron, with the lime, ahimina, and other bases, combines with the silica to form a slag, fluid at the high temperature prevailing at the tuyeres.
The molten slag and matte flow from the furnace to the fore-hearth where the separation is effected, and the supernatant da^, freed from matte, escapes by an overflow spout, and is received into slag-pots and conveyed to the dump, see Fig. 208. Another means of disposing of the slag is to allow the stream of slag from the fore-hearth to fall into a launder and be caught by a horizontal jet of water to break it into drops and cool it in granules about wheat-size. The granules are carried away by the water, in a cast-iron lined launder to the dump. The matte is
FiQ. 208.— Pouring Slag.
tapped from the fore-hearth as it acciunulates, through a tap-hole near the bottom, and flows over the matte spout shown at the right of the transverse section, Fig. 195.
Copier Matte
Matte is an artificial sulphide formed in smelting as a result of the union of sulphur with bases. Iron sulphide (FeS), such as is used in the making of hydr(^n sulphide in the laboratory, is the simplest form. To produce it in small quantities, a covered assay crucible may be filled with shingle-nails and brought to a white heat in a wind-furnace and roll-sulphur added gradually until the content fuses. The sulphide is then poured, and broken for use. In smelting a charge containing sulphiu-, si^rap-iron will take up the sulphur and form matte. If copper oxide or copper sulphide is present in the charge, the sulphur takra the copper to form the matte in preference to taking the iron. When the copper is exhausted, the excess of sulphur expends any combining power that may remain by taking iron. We thus have a copper-iron sulphide, called copper matte. If lead or nickel are present in the chai^, they partly enter the matte. Magnetic iron oxide, taken from the charge, also enters the matte. Thus we get, finally, a complex compound, as the following table shows:
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Blast-Furnace Smelting Of Sulphide Ores
coHPOsmon op copraiR hatte
P-'Si..
8p.0r.
Sudbury nvatte contains abo 15.56 per cent nickel lepbtcing copper. It will be noted that the percentage of sulphur (23 to 24} is approximately the same in all cases.
Melting-point of Mattes.— 32.6 per cent, 87S° C; 40.7 per cent, 955°; 61.2 per cent. 1070°; 71.1 per cent (whitemetal), 1121°, 6 OusSFeS; 80.1 percent (pimple metal), 1098° remaining after separating bottoms; metallic copper, 1083°. Bottoms contain Cu, 60,per cent, Pb 33 per orait.
Coppbr-Fohhace Slags
Variation in slag composition is permismble in copper-smelting, the requirement being that the slag be flxiid to Sow from the tap-hote of the furnace. Slags having the maximum content of silica and of bases, as Bhown below, are employed successfully in the blast-furnace.
In silver-lead smelting practice, such variations are not allowaUe. Sb^ varying from the composition found in practice to be satisfactory, even though they be fluid and run well, carry off both lead and silver. Id copper practice such slags would be clean and free from copper.
THE COMPOSinon OP SLAOS
pj&..
P.'i'gi.
Minimum
Maximum
A slag low in silica could not carry much of the alkaline-earth bases, and would be high in iron, and of a specific gravity over 3.7. A siHcioua slag would work well with a heavy limy base, and would have a specific gravity of 3.5. Since the separation of slag from matte results'&om the difFerenoe in specific gravity of the two substances, we expect a better separation the lighter and mote dlicioufi the slags. Matte takes up zinc sulphide, where much is prosent, and becomes lifter, so in this req>ect, zinc is detrimental to effective separation.
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Calctjlation Of Charge 375
Calculation Op Charob Foe Matte Sueltdig
A low sulphur charge may consut of roasted ore, oxidized coppeT>ote and siliciouB ores containing gold and diver. The lequiiement is that the charge contain copper-bearing ore, and enough sulphur to form with the copper a suitable matte that will take up the gold and silver that the charge contains. Enough flux is added to the chaige to make a suitable slag, and 10 to 15 per cent coke or charcoal to smelt the mixture.
The products from the furnace are slag and matte, the former being the result of the imion of the sihca of the charge and the fuel, with the bases that are present. A part of tbe sulphiu' in the charge is volatilised by the heat of the furnace, but a large part, stall remaining, combines with the copper and a part of the iron, to form tbe complex sulphide called matte. The iron not needed for the matte enters tbe slag. Since the copper-furnace sl^ woiy vary within wide limits, we use a ffllicious ore where silica is abundant and a basic one where plenty of iron is pieeent, or in treating basic ores.
H.6.
"■•"■
Cu-
BiC.
Fs+Mo.
Ca0+M«0
S.
Naow of On.
WBt.
Dry.
Per
Cnt.
Wt.
&
Wt.
&.
wt.
&
wt.
Wt.
0
1
i'.i
'it
Im
jo.o
Coin.
Im-
4
?wm
tw-fd
sa
Cu and F< in matte -
M
Foralac-
Its
75
?tiam
tf-
Cu
and F
Inn
■"•-
31fi
FeO+CaO - M per « Other baaea - 10 per « 100 per « FeOH-CaO 5A
Co +Fe Cu+Fe TO
BiO.
3S
Above is given a charge calculation, in which the problem is to treat a single roasted ore, producing a slag of predetermined composition, and a matte that will take up the copper that is present. Limestone is tbe only fhix to be used. The charge is of a size to M the charge-car or buggy in which it is brought to the furnace.
We will adopt 1000 lb. as a weight of the roasted ore, having the composition Cu 10 per cent, Si02 25 per cent, Fe 30 per cent, and roasted so
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376 BLAST-FURNACE SMELTINa OF SULPHIDE ORES
that 10 per cent sulphur remains. This is to be smelted with limestone containing Si02 4 per cent and CsO 52 per cent to produce a sla^ of SiOi 35 per cent and bases (FeO and CaO) 55 per cent, together 90 per cent, leaving 10 per cent to allow for other elements. The slf^ has been chosm of this composition as one that has been found to work well. The coke has 12 per cent ash that consistB of SiOj 60 per cent, Fe 10 per cent, and CaO 15 per cent. These figures, ctdculated to the coke, are SiOa 7.2 per cent, Fe 1.2 per cent, and CaO 1,8 per cent,
A metaltui^Bt, accustomed to types of ore, knows approximately how much flux he needs. Suppose we decide upon 300 lb. flux. For the calculation we enter on the charge sheet the 1000 lb. ore, the 300 lb. limeetone, and 10 per cent of these or 130 lb. coke in the column of dry wei^t. When the exact figures have been computed, the wet wei^ts may be inserted in the adjoining column, using the figures for per cent given in the column marked HzO. The percentage of ore, flux, and fuel aie tben written in the appropriate columns, and the corresponding weights, calculated to the nearest pound, are written in and the totals added.
Beneath, and at the left of the sheet, tabulate the slag composition, lilnd the ratio of base to silica, which in this case will be 1.57 to 1. On the right of the sheet write the matte composition. We know it will carry 23 per cent sulphur, and roughly 69 per cent copper and iron. Also find the ratio of sulphur to base, which here is 1 to 3, or the factor 3.
Let us first consider the sulphiu'. Experience shows that in regular matte smelting we can depend upon a loss by volatilization of 20 to 40 per cent sulphur. We take 25 per cent as an average, and thus 75 per cent of the sulphur is left to form matte. This is 75 lb., and multiplied by the factor 3 indicates that 225 lb. Cu and Fe together are needed te satisfy the sulphur.
The slag produced is calculated by dividing the weight by the per cent of silica, expressed decimally, or 271-5-0.35 = 770 lb. Allowing 0.5 percent copper for the slag (and in good work it should not exceed this, the weight BO lost is 4 lb., leaving 96 lb. to enter the matte. Subtracting this weight of copper from the total 225 lb. of copper and iron together needed for the matte, we get 129 lb. iron entering the matte, out of the total 302 lb. in the charge. The remainder (173 lb.) is available for the slag. But the iron existing in the charge as ferric iron is reduced to ferrous form, and we have, in the ratio of atomic weights, 56 parts Fe equal 72 of FeO, or 173 lb. Fe equal 222 lb. FeO. To this we add the 158 lb. CaO, making 380 lb. (rf the two bases. Multiplying the silica (271 lb.) by the factor 1.57 we find we need 425 lb. of the bases FeO and CaO, so that we have a deficit of 45 lb. Now, since the limestone consists approximately half of CaO, we need to add 90 lb. limestone to the charge, making in all 390 lb. as the required amount. Erase where needed, and re-calculate the char^-
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Pyrite Matte Smelting 377
throi^hout. This time we E^ould come within a few poimda of the correct amount. As long as it is within 10 lb. it is close enough, since variations in the ores, imperfect weighing, and variation in the amount of sulphur volatilized easily exceeds such differences. When, by experience, we have learned the actual percent^© of volatilization we substitute it for that above assumed. The actual percentage of copper and iron in the matte is taken in the same way.
The grade of the matte in copper is learned from the ratio of the sulphur (75 lb.) to the copper (96 lb.) or 23 to 29 per cent. In the same way we compute from the respective weights the percentage of SiOz, FeO and CaOi.their afg^gate being 90 per cent.
The metallurgist seldom can count on the sl^ and matte coming from the furnace precisely as calculated. There is a little variation due to the causes already mentioned. When the slag from a newly calculated charge comes down, a sample should be taken and a rapid detennination made for Cu, SiOj, FeO, and CaO. As an approximate rule, to increase an ingredient of the charge a given percentage, add to it the fractional part expressed by its ratio to the remainder of the 100 per cent. Thus, if analysis gives 33 per cent and we wish to increase it to 35 per cent, then to the ^ of 271 =-16.4 lb. add H^ (i) of 16.4 lb., making the total silica to be added 25 lb.
Pyiute Hattb Shblting
This consists in treating in a blast-furnace, such as that shown in Tig, 196, sulphide ore consisting largely of pyrite and chalcopyrite. The ore carries gold and silver, which are recovered in the copper-bearing matte produced. No preliminary roasting is given the ore, and the smelting is conducted m such a way that 70 to 80 per cent of the sulphur is burned in the furnace while the remainder, uniting with iron and the copper, forms the matte which acts as collector for the gold and silver. A slag is formed from the silica of the gangue and the bases of the ore and flux. At times when the quantity of base, especially iron, is large it is necessary, in order to make a suitable slag, to add silicious ore. The matte and slag flowing together from the furnace separate in the fore-hearth.
It will be noticed that the slow and expensive preliminary roasting of the sulphide ores is obviated, and that the amount of fuel needed is small (1.5 to 6 per cent) because of the heat developed by the burning of the sulphide. Pyrite or chalcopyrite contains iron that is available both for matte and slag, and when the matte can spare it for the slag the iron serves to flux the silica of iron-free ores on the charge. Iron ore or limestone acts in the same way, and either of them, though generally the latter, may be added for the purpose.
An iron jnstte alone does not entirely collect the gold and silver from
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378 Blast-Furnace Bmeltino Of Sulphide Ores
the ore-charge, and it has been found that copper, to the extent of 0.5 per cent or mon, should be present to insure the collection of theee metak in the matte. The alag then will be neaiiy free from the pieciow metals. Copper, thei^oie, acta as an efficient collector.
As a result of burning 70 to 80 per cent (tf the sulphur of tibe charge, there remains only 30 to 20 per cent to form matte. The remainii^ sulphur first takes up copper, for which it has a greater affinity than for iron. It is the burning off of the large amount of sulphur that enables one to di^sense with roasting and to diminish the amount of matte produced. The matte produced per ton of ore, or the matte-fall, may be expressed as a percentage, or as a concentration of so many tons into one of matte. Thus, witii a production of 200 lb. matte per ton of ore, we have a 10 per cent matte^all, or a concentration of 10 into 1. It is desirable to concentrate the ore into a onall bulk of matte. To show how much concentration is effected, both in regular matte smelting and in pyrite smelting, we enter upon the following cmsiderations:
In regular smelting (with a chai^ containing 8 per cent sulphur, the volatilization-losB being 25 per cent, and the matte to contain 25 per cent sulphur), we have from 100 lb. ore 75 per cent of 8 per cent = 6 lb. Bul|dnir to form matte. This makes 24 lb. matte and nsults in a concentration oi 4.2 into 1.
In pyrite smelting with a cha^e containit^ 30 per cent suli^iur, the volatilisation loes being 80 per cent and the matte still to contain 25 per cent sulphur, we have from 100 lb. of ore 20 per cent of 30 per cent=6lb. of sulphur to form matte. This makes 24 lb. of matte, the atune ctmcentration as in the regular matte smelting just specified.
It will be noted that the percent^ of volatilisati<Ri, or the amount ot sulphur burned, varies with the chai^. It is low when only roasted or oxidised oice are used, and hi^ for raw or unioasted ores, especially those containing pyrite. A d^ect inherent m pyrite smelting is liie difficult of regulating this loss. When the furnace, which is burning the ri^t amount of sulphur, begins to run slow from tmy cause, the volatilization may increase to the point <rf bumii^ the entire content of sulphur, so that no matte is produced. On the other hand, when the furnace begs to run fast, much matte, low in copper, is produced. The principal difficult in pyrite smelting is the regulatioii of the matte-fall.
REACTIONS m PTRITE MATTE SBCBLTQTO
Fig. 209 is a croes-eection of a matting blast furnace smelting a pyritic charge, and consisting of sulphide ores with the addition of enou^ quatti or silicioua ore to make a suitable slag. With this charge 3 per cent of coke is used, but this does not inteifeie with the reactions.
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Reaction In Pyrite Smelting
At the surface of the charge, which is mamtamed at 12 ft. above the tuyeres, where the temperature may be at 250° C, the heat drives off from the FeSs a portion of its sulphur, leaving as FesS*. The eacapii^ sulphur fume, encountermg the air entering the feed-door, bums with the ohaiacteriatic bhie flame to SOa-
By the time the charge has gone down 5 ft. in the furnace where the temperature is much higher, more sulphur has been expelled, leaving FeS.
At 7 ft. down, luid at a temperature of 925 to 950° C, diseodation continuing, we find Fe5S(=4FeS+Fe, or a condition in which four equivalents of FeS hold one of Fe in solution, so that, if a sample of the compound in a molten condition could be withdrawn from the furnace we would find iron separating from the iron sulphide on cooling.
The FesSi at this zone b^pns to melt and falls, entering the silicious, porous structure or " nucleus " shown above and below the tuyeres in F^. 209.
Within this porous structure the downward trickling sulphide is encountering the rifang air according to the reaction
Fta. aOQ.-'Section of Furnace in I^ritie Smelting.
= 5FeO + 4S0a 5X66,400 4X71,000 - 520,800
This is equivalent to 1874 pound-calories per pound of iron present.
The forming I^^O at once unites with the sihca present to form a molten slag which continues its journey to the crucible. The rising gases, intensely heated as the result of the reactions, expand, keei»ng the nucleus porous. The ore-column, therefore, instead of resting on a bed of burning coke in the crucible as in regular matte smelting, is sustained upon a network of quarts pieces constituting the nuclexis, and which have thus far escaped slagging. The nucleus extends from the crucible upward to the sone of fumcm of the iron sulphide 5 ft. above the tuyeres, having its greatest development at or about the tuyeres. At the sides of the furnaces the network is moving slowly downward, and is sustaining the portion which is descending tegulariy.
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380 Blast-Furnace Smelting Op Sulphide Ores
Generally not all the FeS is oxidised, but a part, together with any copper sulphide present, foimfl matte and in molten cimdition seeks the crucible.
In a regularly working furnace produ<!ing its eq;uivalent slag, this slag is, within limita, not changed in composition by a change in the amount of silica. It must be understood that the addition of silica has the effect of increasing the degree of concentration, that is, of raising the grade oi the matte. There is then more iron oxidized and slagged off, but the effect is not to make the slag more irony, but simply to increase its quantity. If too much mliciouB ore is added, the excess remains undigested and chokes the furnace. If too little is added, the amount of matte increases while its tenor in copper becomes less, and at the same time the amount tA dag decieaaee, but without altering its composition. Considered in anotb^ way, the pyrite furnace chooses its own dag.
As compared with other kinds of smeltmg, an abundance of air should be supplied to the furnace, not less than 300 cu. ft. to 1 lb. of sulphur present in the chai^.
At Mt. Lyell, Tasmania, where great success has been attained in pyritic smelting, two parts of heavy sulphide ore containing 2 to 2.25 per cent Cu are used to one of silicious ore of 70 per cent SiOs with a concentration of 18 to 20 parts into one, producing a matte containing 40 per cent Cu. Of the large amount of iron present 95 per cent is burnt or oxidized, the small remainder going into the matte. Under given conditions, a rise from 95 to 96 per cent of iron oxidized results in an increase in the grade of the matte to 50 per cent Cu.
The porous condition of the nucleus is practically preserved as the result of the reactions there taking place, while it is the duty of the metallurgist to see that such loose and porous condition is suitably maintained elite-where in the shaft of the furnace, which must be kept properly open both below and above this fiery net-work. The proper maintenance of this condition is one of the principal secrets of success in order to avoid freeieups and to use the minimiun amount of coke.
Pyrite Smelting in Two Stages. — For low-grade ores, carrying 2 per cent copper, for example, a concentration of ten into one gives matte of 20 per cent copper. By roasting the matte, or by smelting it pyritically, it is possible to increase the grade to 40 per cent or more, and this product can be treated in the copper-converter and brought to the grade of Uistercopper. An ore containing 5 per cent copper can be smelted te give a matte of 40 per cent copper, so that the second smelting with the additional expense can be omitted. Now while it would not pay to smelt a copper ore of a grade as low as 2 per cent copper, for the copper alone, if the ore contained gold and silver the recovery of these metals would justify the expense of smelting.
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Calculation Of Chakge
Two-Stage Smelting at Ducktown, Tenn. — At the works of the Tennessee Copper Co., the process is a two-stage one, the ore being smelted to give matte of 10 per cent Cu (" ore-smeltiiig "), this matte being re-tiBated in another furnace to produce a 36 to 45 per cent matte (" matte-smelting "). The slag from the second furnace, not yet sufficiently clean to throw away, is lemelted in the first furnace. The second matte is then converted to blister-copper of 98 per cent.
The furnaces (of the type shown in Fig. 196) are 196 in. long by 56 in. wide at the tuyere level, and have lai^ circular fore-bearths, 16 ft. outside diameter by 5 ft. deep, lined with chromite brick to resist the corrosive action of the low-^rade matte produced. The first, or ore-furnace, treats 400 tons of charge daily, with a coke-coneumption of 2.1 per cent. The second or concentration furnace smelts 280 to 300 tons of matte, using 3.5 per cent coke. The following chai^-sheet gives details regarding the charge and the matte produced in both the stages, and shows how such charges are cconputed in pyritic smelting.
CALCULATIOIf OF CHARGE IS PTKITE SMELTHfG
First Stage, the Ore Charge.— In these calculations the quantity of coke is so small that no computation is required for the ash. The quantity of base in the ore is so large, and the silica is so low that it has been necessary to add siliciouB material (in this case quartz-rock) to the charge in order to obtain a slag of 35 per cent silica. The problem is to compute the amount of quartz to be added to ^ve such a slag.
Charge Sheet I
Ore-Smelting Furnace.
w„„,.
H.O-
Cu.
SiO.,
F. + Mn,
CO+MgO
s.
Name ol Or»
Dry.
Per
Cent.
"■■
C*!"!
L..
cS.
Li,
Per Cent
Ce"t. ">■
Cok.
Cuin
nSUi matte
B7;o
nij
1
In
U
aeo
FeO+CaO-SO Operc
Fb+Cu- 8-: Cu+F«
We represent on the charge sheet the ores that are to be run, using amounts in accord with the rate at which the respective ores are supplied
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382 Blast-Furnace Smelting Of Sulphide Ores
(1000 lb. Polk county OTe and 3000 lb. Bunra-Burra ore). Experience shows titan for such a charge and for the quantity of ferrous iron and lime present, we may enter the quantity of ailicious material as 700 lb. We use a dag of 35 per cent SiOa and dd per cent FeO and CaO, making in all 00 per cent. With this slag, experience shows we may figure on 1.3 per cent sulphur and 0.2 per cent copper for this low^rade matte. A little zinc, when that element is present in the charge, also enters the slag. This is shown to be 0.3 per cent.
The percentage of ingi^ients of the chai^ is written and carried out in the leepective columns, and the coltmrns are added.
Beginning with the sulphur, of the 1 1 13 lb. present, we have:
Lb.
Sulphur left for the matte 180
The matte is assumed to contain 25 per cent sulphur, 65 per cent copper and iron, and 1.7 per cent zinc. We estimate that 80 per cent of the sine will be volatilized. It is understood that in smelting otiier ores than these, the actual quantities of the different elements in the matte and sbg will be determined and those figures substituted for the ones above.
Of the copper, 0.2 per oent of 3300 lb. or 6 lb. goes into the slag, leaving 81 lb. for the matte. Multiplying the sulphur for ihe matte, 180 lb., by the factor 2.6 we get the total Fe and On needed for the matte, 468 lb. ; subtracting Cu for matte, 81 lb. ; leaves Fe for matte, 387 lb.
But the total iron in the charge is 1482 lb., so that we have :
Lb.
Fe in matte 387
Fe left for slag 1095
Total 1482
The iron in the slag occiirs as FK), so that we must take 1408 lb. FeO equivalent to the 1095 lb. Fe. Adding to this the CaO, 341 lb., we get P^O+CaO-1749 lb., which multifJied by the factor 0.636, gives SiOaa 1112 lb.
Lb.
Actual ailica in charge 1168
Silica needed 1112
Baica in excess fi6
By erasing the trial amount 700 lb. of qxiartz, and substituting 650 lb. then recalculating the charge, we get an approximation within 10 to 20 lb., which is accurate enough for practical purposes. The percent^e of copper
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Concentration Of Matte
in the matte is computed according to the proportion 159 : 81 :: 25 per cent : 10,4 per cent.
Second Stage, Matte Concentration. — ^This is run witli slag from the converting operation, Mid quartz ore in sufficient quantity to produce a slag of the same composition as that of the ore-charge, except that it has 1 per cent of sulphur and 0.7 per cent copper.
CHARaB SH
BT ri
Matte-Concentration Fuknacb
Wwoh™.
Cu.
BiO..
Fb+Md.
s.
Nune of Ore.
W«t,
Drr.
&.
„
cSS.
Lh.
p.,
Cent
Lb,
&
c£;.
Lb.
1
c.",:':
Ss:
'it
lias i4se
„:o
1S70
as,o
lai
%Oi-35.0 par cent. Ps-fCu-e6i>iT0«nt.
FeO +C*0 -S5.0 per cent. B -2S par «Dt.
S-l,Of«r«i.l. Cu+F«
SliD -1480 +0.3S -4300 lb. Futor^ -0.636.
The charge is estimated as in " Charge Sheet I," with a volatilization of 70 per cent of the sulphur, as experience has shown the result commonly to be. In the matte we have:
Lb. Par Cant.
S 181-25.0
Cu 278=83.4
652 = 90.0 Proceeding with the calculation for the quartz we have:
Lb.
Total iron 1870
Iron in matt« 193
Iron for elag 1677
or FeO=2157, and the total base is 2266 lb.
This gives the sihca needed, 2266X0.636 = 1509 lb. But we have already 1486 lb., and the difference may be made up by increasing the quartz 20 lb.
Concentration of the Matte. — In an example above, we obtained a matte 10 per cent in copper^ and with copper low in the chai^, the per-
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384 Blast-Furnace Smelting Of Sulphide Ores
cetitage may be even less. The matte is of too low grade to ship away, or to bring to the grade of bhster-copper in a converter. It must be concentrated to one of higher grade. If we wero to heap-roast the matte and then melt it with silicious ore in a blast-furnace, we should obtain a small quantity of matte of a high grade.
There is, however, the expense and delay of the roasting to consider, and it has been sought to smelt the matte raw, with silicious ore, with the idea of burning off the sulphur in the blast-furnace. In tegular matte-smelting, were this attempted, the matte would run through little diminished in quantity and little changed in grade, but by the new method, using little fuel, an abundant blast, and silicious slag, the concentration can be obtained.
Disposal Of The Slag
The sIe^ from a blast-furnace, being a waste material, is disposed of in the cheapest way possible. In the case of small furnaces, as it flows
Fi<i. 210.— Electric Trolley .System (retii.ivinK slaR-pot).
from the fore-hearth, it is caught in wheeled slag-pots (slag-carts), Fig. 200, that are taken to the edge of the slag-dump when filled and poured. As the dump grows the expense increases, and large slag cars, Fig. 210, are used. The cars are moved either by hoTses, or an industrial locomotive, or by trolley.
Another cheap and favorite way is to granulate the sli^. To do this a cast^nm launder is arranged to receive the slag as it falls from the spout
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Blast-Furnace Vs. Reverberatisry Smelting 385
of the foie-beartb. The latmder has a grade of 1 ia. to the foot, and through it water is made to flow constantly. In addition, a horizontal fiattened jet of water strikes the falling slag, instantly cooling and breaking it into granules of various sizes averaging A in-in diameter. The flow of water carries the sl^ to the dump.
BLAST-FUIUTACB VS. BKVERBERATORT SHBLTUrO
Predictions have recently been made that the reverberatory was bound to supplant the blast-furnace, because of the advantage the former pessesaed in the treatment of the finer ore and flotation concentrates. It is pointed out, however, that where ore is coarse, and where it is possible to avoid roasting, the blast-furnace has its advantages even for sulphide ores.
" There ia no doubt that at the moment, in favored localities where pulverized coal or fuel-oil can be obtained at a much cheaper rate than coke, the reverberatory has the better of the-argument, but there is always something turning up in favor of the other side in every controversy. It resembles the perpetual iight between armament and projectiles that we are all much more familiar with at present than we were prior to 1914.
"Both styles of fiimaces have their field, but for the moment, owing to the great increase in tonnage treated by oil-flotation and the improvements in reverberatory practice, the reverberatory seems to be gainii^ materially in tonnage treated.
"The latest improvement in blast-furnace practice ia the intreduction of pulverized coal at the tuyeres, the notable examples of this being the Tennessee Copper Co. smelter and the International Nickel Co.'s plant at Copper Cliff, Ont. If the experiments now being tried at various plants, besides the two mentioned, prove that pulverized coal can be used economically in the blast furnace, either with or without a proportion of coke, the probabihties being that a certain proportion of coke will be necessary, the existing interest in blast-furnace practice will receive an impetus.
"The new plant being erected in Chile for the Braden Copper Co. contemplates the use of the blast-furnaces entirely, using nodulizing-furnaces to prepare the chaige. To-day it is about a stand-off in cost between nodulizing or sintering for blast-furnace practice and roasting for reverberatory practice."*
Methods of Reverberatory Smelting. — The essentials for successful reverberatory practice are self-fluxing ores, cheap fuel, and cheap silica brick, whereas for blast-furnace practice the charge may be more refractory, but must be either sintered, nodidized or come naturally in lumps, and in addition the fuel and power must be comparatively cheap.
While the blast-furnace in general is the cheapest means of smelting coppei^bearing ores in coarse or liunp form, one objection to it is that the • E. P. Mathewten.
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386 Blast-Furnace Smeltino Of Sulphide Ores
blast may cany away 5 to 10 per cent of the fine dusty oie. This may be settled as fiue-dust in flue or dust chambers, made into briquettes and resmelted, but the additional expense should be avoided if possible. Ore or concentrate in fine condition is better treated in the quieter atmosphere of the reverberatory furnace. If raw ore is treated in such a furnace about 25 per cent of the sulphur only would be expelled; hence, before smelting, sulphide ore would be roasted.
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Chapter Xxx
Rbvesberatory Smelting
Two methods have been evolved for the reverberatory smelting of copper ores, vis., the Welsh process and the reverberatory matte smelting process. The Welsh process possesses the advantage that it can be used on a great variety of ores, the final product beii^ a blister copper. In reverberatory matte smelting a great tonnage of roasted ore is put through, and the resulting product, which b, in the form of copper matte, must imdergo a further treatment in the converter to bring it to the stage of blister copper.
This coDfflste in treatii^ copper ore (sulphide and oxide as well as silicious ore) hy a series of roastings and fusions to raise the grade of the copper in the product finally to blister copper, which is subsequently refined electrolytically, as any blister copper containing gold and silver naturally would be. The process has the advantage that a variety of ores, p,^ 211. -Reverberatory Smelting Furnace, both coarse and fine, can be treated in a few small furnaces with a small investment of plant.
Fig, 211 is a furnace lined with refractory material having a shallow basin-diaped hearth. Square cross-bars in the fire at a sustain the grate
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bM8 of 1) in. square iron (not shown). A deep fire ia maintained to the top of the fire door, e, and aome cinder is allowed to accumulate on the grate, but so as to keep the fire open. Any grate bar can be [lUshed aside and the cinder dropped into the ash pit h when cleaning the fire. Next to the firebox is a " bridge " 2 ft, 6 in. wide, to confine the fuel to the firebox. A brick arch beneath it sustains the hearth foundation of brasque and on this the refractory brick of the hearth. The charge is thrown in at the rade door, which is then quickly bricked up. The furnace is Btin^ from the front door, and the sl^ and matte removed by rabble. The producte (A combustion escape by a port or opening in the roof and l^ a Sue are carried to the stack or chimney. The part of the roof toward the outlet port is called the " verb."
Smelting Operations Bt The Welsh Process
We may divide the emelting operation into five parts:
(1) " Caldnlng " the Ore. — Sulphide ore containing 5 to 15 per cent copper is roasted in a hand-reverberatory roaster (see Fig. 29) until not more than 5 per cent sulphur is left.
(2) Fusion erf Ore. — The roasted ore is charged in a reverberatory furnace with such oxidized copper ore as is available, and melted. The sulphur contained in the roasted ore, with the copper and some of the iron, forms a matte of 35 per cent copper, called " coarse metal." The dlica, uniting with the ferrous oxide not taken by the matte, and also with the alumina, and the alkaline-earth bases, forms a slag, fusible at the high temperature of the furnace. The molten bath boils from the escape of trioxidc resulting from the reaction of the ferric iron upon unroasted ferrous sulphide, or from the decomposition of barium, lead, or zinc sulph&te by silica, thus:
The sulphuric anhydride escapes as a gas, and upon meeting the moisture of the air at the top of the stack, changes to a white fimie of H2SO4.
(3) Calciniog Coarse Metal. — ^The coarse metal or matte that is run from the reverberatory furnace in operation (2) into sand beds, is crushed to pass a 5-mesh screen and fed to another hand-roaster. Rich sulphide of 20 to 70 per cent copper also is crushed and added to the charge. The whole is roasted until it contains not more than 5 per cent sulphur.
(4) Second Reverberator; Fusion. — The roasted material, now of 35 to 50 per cent copper, is charged into a fusion-furnace with oxidized ores containing 20 to 70 per cent copper. When the charge is melted
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The Wei£H Process 389
there results a matte of 75 per cent copper, called " white metal," composed chiefly of copper sulphide. As before, the dlica contaised in the ore added to the cbai^ uoitee with the ferrous itoq and other bases to fonn dag. This dag, however, having been made from such rich material, contains much copper and is not to be thrown away but returned to another charge in the fusion-furnace of operation (2).
(5) " Roasting " and Formation of Blister-copper. — The white metal is charged in lai^ pieoee, as Iraoken when removing from the sand molds, into a reverberatory fusion-fumace where it is piled m an open fashion, particularly near the bridge. It is fired gradually for several hours with an oxidizing flame. A supply of air is admitted at a ntunber of ports or openii^s 2.5 in. square in the roof over the fire-bridge, and at the sides of the f oroace near the bri<^. The operation is called " roasting." The oxidizii^ flame, acting at the surface of the lumps and upon the drops trickling down, converts a portion into cuprous oxide, so that we have present copper both as sulphide and as oxide. Finally the heat is raised and the whole charge is melted down, according to the reaction;
2X42,000 20,200 71,000 --33,200.
Copious fumes of SOa issue from the boiling surface of the molten charge. Slag rich in copper is produced, getting the sihca partly from the interior walls of the furnace, partly from silicious but entirely oxidized ore that has been added to supply silica. The slag is returned to operation (4) . Finally the blister-copper is tapped. The metal obtains this name from the fact that, upon cooling, occluded gas seeking to escape from the molten metal, forms blisters on the surface of the pigs of metal.
The blister-copper now contains 98 per cent copper, but also impurities that must be removed to make it suitable for market. The refinii^ process is described elsewhere. In case the copper contains gold and silver, taken frcon the ores that supplied the copper, it is customary to remelt it, pole it to remove copper oxide, and to cast it into anodes for electrolytic refining.
Treatments of &e Bottoms. — Impure bottoms are remelted and cast into anodes for electrolytic refining, in which the impurities and gold are separated from the copper; or they may be formed into an inferior grade of copper (castii^-copper) as follows: A chai^ is put into the blister-furnace as in operation (5), consisting of 14,000 lb. of 75 per cent roasted white metal, 21,000 lb. raw white metal, 8000 lb. bottoms and 1000 lb. silicious ore. This is melted down, and then is added 6000 lb. more roasted white metal. The charge aggregates 50,000 lb. This is treated precisely like the regular b)ister-chai^, but it yields a higher percentage of copper.
Reverberatory Smelting
THE DIRECT PROCESS OF REVERBBRATORT SHELTinG
This is a modification of the Wel^ method of producing blister^ copper. Instead of " roasting " the matte or white metal in himp fonn in the blister-furnace or process (S), a portion is ground to S-mesh size and calcined or roasted in a separate roasting-fumace, of either the hand or the mechanical type.
Into a melting-furnace, called the " bUster-fumaoe," is charged 14,000 lb. of the roasted white metal, still retidnii^ 4 to 6 per cent sulphur, 3500 lb. raw unroasted white metal, 4000 to 8000 lb. alag from a former charge, and 60O lb. eihcious ore to unite with the FeO and other base present. When this has been melted, 6000 lb. more roasted matte is added, maMng a total of 23,500 lb. matte chai^^. When all is fused, the reaction b^^ns. The surface of the charge is seen to be seething and boiling, and esc^'ng bubbles of gas are set fiee according to the following reaction:
The bath is then dammed to remove the slag, which contains copper oxide, deserved in part for the next charge and in part sent back to stage (4) of the Welsh process. From the charge bere specified there are produced 75 pigs weighing 230 lb. each, or 17,250 lb. bhster-copper, and alao 23 pots of slag weighing 400 lb. or 9200 lb. containing 12 to 15 per cent copper. Slag is removed several times during the period. The copper, when free from slag, is tapped into a defining-fumace at a lower level. The refining is done in a 14-by 22-ft. furnace carrying a deep chai^ of copper. This chai^ is made from copper scrap, high-grade " mineral " of over 80 per cent copper, and mass-copper. It is melted, and a tJistercopper, comparatively free from impurities, is obtained.
Large-Scale Reverbbratort Hattb Shelting
This and blast-furnace smelting are practically the two methods tiS treating copper ores m the United States. Such ores, principally sulphides, are roasted and smelted in large reverberatory furnaces for the production of matte and of a slag low in copper, which is sent to waste. Fuel, rapidly burned, maintains the furnaces at a temperature above the smelting [mint of the forming slag. We may divide them according to the method of heating into (1) direct-fired furnaces having a firebox for coal bumit^; (2) furnaces fired witii pulverized coal, and (3) oil-fired fumaceB. Except at the fire-end they are essentially the same.
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(1) THE DIRECT COAL-FIRED FOSlfACB
We give at Fig. 212, in plan and elevation, & large furnace showing, at the " back end," the ash pit and the large grate, 8 by 16 ft. area, and the solid fire-bridge immediately in front of it. At the outlet end the fhie comes out at the roof. There is ample room in front for Bkinmiing and tapping the slag. This runs into a water-filled bosh where it is granulated and swept away by a powerful horizontal jet of water. At two points on the aide the matte is tapped oS at the hearth level and runs by gravity along a matte launder to a matte ladle set at a low level to receive it. At the left are shown the waste-heat Sterling boilers of 300 H.P. each. Thence the gases, having given up much of their heat pass on by an under-
Fio. 212. — Coal-liked Reverberatory Fumace.
ground flue to the main stack. In case it is desired to cut out the boilers for re|^r the damper to the branch " rnideiground flue " is opened and the boiler damper is closed. Two boilers are heated by the waste gas from the fumace, and develop, together, 600 H.P. The furnace treats, on an average, 275 tons in twenty-four houra, producing a 40 per cent matte concentrating 4 into 1.
The charge con^ste of hot roasted ore (" calcines ") from MacDougall roasters. Fig. 77, and by analysis is shown to be composed as follows: Cu 9 per cent, FeO 24.4 per cent, CaO 2.9 per cent, S 8 per cent, SiOa 26 per cent. Elvery eighty minutes a chai^ of 15 tons is dropped into the fumace near the fire-bridge. This falls upon the bath of molten matte and slag that the fumace contains. It spreads in all directions, and much of it floats gradually toward the front. It readily melts by contact with the molten slag and matte below and the flame above. In
this great reservoir of heat there is but little variation in temperature, and the flame is transparent.
Ewry four hours 45 to SO tons of slag is removed in fifteen minutes from the furnace and allowed to flow from the front door in a thick stream. It is granulated by a strong horizontal stream of water as it falls into the waste-launder. The water sweeps it away to the dump several hundred yards from the furnace. The matte is kept at a nearly uniform level, 10 tons being tapped out at a time, while the total amount in the furnace is 100 to 150 tons.
The action of the slag upon the furnace is to erode or scour it, but because of the width, the sides are less acted upon than would be the case in a narrow furnace. To repair the furnace, both slag and matte are drawn off completely, then the side-doors are opened, and sand is thrown across the furnace against the sides where eaten away by the dag. They thus are protected against the inroads of the slag. A furnace runs six or eight months and then has to be shut down for the thorough repair of the roo^ walls, and bridge. These parts are of sihca brick, the walk being 30 in. thick, the roof 15 in. Silica bricks are practically infusible but expand on heating, so that allowance is made by leaving transveiw slits in the roof. These close when the furnace is at full heat. An important point in efficient working is to have the outlet-flue, or " neck," of the proper size. It must be large-to insure good draft, and yet retain the flame in the furnace. In the furnace shown it is 60 by 38 in. or 16 sq. ft. area.
The temperature at fire-bridge is 1550° C; at the flue end 1200° C. The slag leaves the furnace at 1120° C. and drops to 1060° at the ov&- flow spout of the settler. Gases reach boiler at 050° and leave at 330° C.
Hat Biluioa rf aboTc Pumue.
Heat in alag 16.2
Heat in matte 3,2
Heat lost in radiation 11.6
Heat loBt in cooling-bridge plate 0.2
Sensible heat in grate droppings 0.8
Heat in steam generated at boileis 32.8
Heat in ga»es passing the boiled 13.2
78.0 The remainiuR 22 per cent must be due to the excess of air over the theoretical amount needed for combustion.
(2) FURNACES FIRBD BY PULVERIZED COAL In Fig. 213 is a sectional plan and elevation of a pulverized coal-fiied furnace, with a hearth 116 ft. long by 19 ft. 9 in. wide and with sidewalk 22) in. thick. At the firing end are two tall charge hoppers, but the moat
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Powdered-Coal Firing 393
of the chai^K ia done by side hoppers having 6-m. feed jApea extending through the roof De&r the side walls. By opening a slide in the feed pipe, calcines can be charged against the wall as fast as the material melts down. At the firing end is an 18-in. blast pipe from two blowers with five air jets
Longitudinal Seijiiun
Fig. 2t3. — Revertxratory Furnace Fired with Pulverized Coal,
Flo. 214. — Method of Charging the Furnace.
to the furnace. From the pulverized coal bin come down five pipes supplying the coal by a screw feed to the tops of the air branches or burners. The coal as it drops into the burner is atomized and blown into the furnace, instantaneously taking fire and filling the furnace with flame. A slag launder at one side of the front end takes the flow from the tap-hole, while
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Reverberatory Smelting
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Oil-Fired Furnaces
a matte-kunder towarde the firing end takes away the matte at a tap-hole set, &8 shown in the sectional elevation at the hearth level. On the opposite mde of the furnace are spare matte launders to be used in case of need. It wil be seen that the furnace bottom is <rf sand fused in layers by heavy firing.
How the furnace is charged is well shown in Fig. 214. The calcine, hot from the roasters, is banked along the wall as at £ and as shown in the typical section of the same figure. There is also a transverse section show-ily the position of the burners. It will be seen from the temperatures given in Fig. 213, plan, that the banking is near the hottest part of the furnace.
SECTIONAL ELEVATION ON CENTER LINE toJ^'It^J^ Flo. 210. — Low-|irpsRure Oil-bumerB. (3) OIL-FIRED FURNACES
Where oil is the cheapest fuel it b to be preferred, and in Fig. 216 are shown in sectional plan and in longitudinal and transverse sections an oil-fired reverberatory furnace. In this case there are six charge hoppers placed at the zone of greatest heat, but no side charging. The chai^ as there melted down flows toward the front end, the matte settling out in quiet. Ore is t^roUKht to these hoppers by a charge car from the roasters. The side walls are thick and sloped up above the slag line at the
side. This furnace has side doors to give access to the hearth and walk for Tepaira. Tapping of alag is done at the front under the outlet flue, and matte just at the point where the furnace begins to nairow. The roof is high toward the firing end, sloping downward gradually to the verb at the front.
There are f o\ir oil burners located well above the slag line and made as shown in Fig. 216. Oil under hi^ pressure is brought by a \-m.. pipe to a burner tip or nozzle, where it is cau^t up and atomized by the air blast, and thus made ready for instant burning with an intense flame.
Burning Temperature <rf Oil in a Reverberatory Furnace. — A fuel oil of the composition C 85.0 per cent, H 12.4 per cent and 3.4 per cent, when completely burned, will yield 10,720 calories per pound. The products of combustion, includii^ 0.5 lb. steam for atomizing the oil, will be (see page 80) 3.11 lb. carbon dioxide; 9X12.4 per cent-0.5 HjO as water, a total of 1.618 lb. and in the steam 1.44 lb., a total of 3.70 lb. oxygen that comes from 13.91 lb. of air containing 10.71 lb. of nitrogen. Hence we have
3 II lb. carbon dioxide @ 0.364-1,132
1.618 lb. water vapor @ 0.77 =1246
The temperature of combustion is therefore ' = 2000° C. nearly.
In practice some excess of air must be used so the temperature is not attEkined. Thus, with the excess of 23 per cent of air, the theoretical temperature may be given at 1800° C. ; and considering cooling influences not more than 1500° to 1600° C. We may assume that the gases lea\'e the furnace at 1200° and the waste heat boilers at 400° C. Due to leakage of air into the furnace and to the cooling influence of the gases escapinj; from the charge the excess gases are increased to, say, 50 per cent. Summarizing we have
Par Cent.
CabriBc value of the fuel 100.0
Distribution of heat:
Taken away by the matte and slag 13.6
Radiation froni the furnace 10.0
Taken up by the boilers .* 34.5
Radiation at the boilers 7,0
Sent to the Stack : 35 0
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Operation Of Rkverberatory Furnace
Opbratiok Of A Large Reverberatort Furnace
These, whether coal fired, pulverized-coal fired or oil fired are charged much in the earoe way. Part of the charge is dropped into the furnace from time to time from hoppers shown near the firing end of the furnace; the rest is added through chai^ tubes so as to nuiintain a bank of ore f^aiust the furnace side wall at the hottest point, thus protecting the wall against the intense heat and the corrosion of molten slag. Ore is brought to the hopper as well as to the side wall hoppers by calcine cars placed on tmnsverae overhead tracks. However, the Cananea oil-fiied furnace is not bo fed. It is provided with side doors through which ore can be thrown for fettling the walls to the opposite aide of the furnace.
Increase of furnace output depends upon the increase of fuel burned with its increased and intenser consequent heat. It has been found that by enlarging the outlet flue to 70 sq. ft. as compared with half that dimension and by increasing the volume of low-pressure atomizing air with the increased fuel, the flame was shorter, began closer to the burner nozzle, and was intenser, so that melting proceeded more rapidly and a tonnage formerly of 400 was increased to 700 tons and over. In one instance in an oil-burning furnace of 100 by 21 ft. hearth area 800 tons was smelted daily, uang 0.622 bbl. of oil (183 lb.) per ton of charge. Even this figure has been improved by charging the calcines promptly from the
In older practice the custom was to remove the slag intermittently, using a rabble to assist the flow. More or less crust and floating material imperfectly smelted was thus withdrawn. In present practice where 400 tons and over is smelted per furnace, the breast has been closed and the tapping holes made of such aperture that the flow is continuous and the crust near it is undisturbed. In this way the copper in the waste slag was cut from 0.4 per cent to 0.3 per cent per ton of slag.
Ribs are now introduced in the roof, arching from the buck-staves between the ribs, thus prolonging the life of the furnace roof.
Reactions And Calculation Of The Charge
The sulphur of the chai^ unites with the copper and iron until its needs are satisfied. The matte thus formed, separating in drops, absorbs the precious metals contained in the ore, and by the greater specific gravity than that of the slag, penetrates downward to the hearth. The silica of the gangue unites with baseH, such as FeO, CaO and AUOg, and forms a fusible slag that floats as a separate layer upon the matte.
When ore is roasted a part of the iron is oxidised to the ferric state.
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and when the charge is fused the ferric iron acts upon the imroasted ferrous BUJphide according to the following equation:
(7) FeS+3Fez03+7Si02 = 7FeSi03+S02.
When ferric iron is present in the ore, sulphur is eliminated often to the extent of 25 to 33 per cent, and we find less matte than would be presait if this reaction did not take place. The iron, thus reduced to ferrous form, enters the slag.
CHARGE CALCULATION FOR A RBVERBERATORT FtnuVACE
Charge Calculation. — ^The charge of a reverberatory furnace |!:eneraUy consists of hot calcmes from the roasting-fumaces to which limestone has been added at the time of roasting.
The charge for the reverberatory may be calculated as in the example under " Regular Matte Smelting." Below is given an example where the composition of the chac^ is known and we desire to compute the consequent composition of the slag and matte. It is assumed that not more than 30 per cent of the sulphur in the roasted ore will be volatilized, that approximately 0.4 per cent copper and 1 per cent sulphur go into the slag, which also carries 90 per cent of the three elements SiOa, FeO, and CaO; that the matte is to contain 90 per cent of the three elements, sulphur, iron, and lime. The ore has been roasted with 10 per cent by weight of crushed limestone. The char^^e-sheet is arranged as in the table, and the pei^ centagee computed and added. The weight of the slag of 35 per cent SiOi would be 1200 lb., containing 1 per cent sulphur, or 12 lb., and this added to the 30 per cent or 42 lb. of the sulphur biimed off, will make 54 Ih., leaving 86 lb. to enter the matte whose weight would be four times this or 344 lb. Some 65 per cent of the matte will be copper and iron. After allowing for 0.4 per cent or 5 lb. as being lost in the ^ag, there remaina 115 lb. of copper, and 108 lb. of iron is consequently needed for the matte. Deducting this from the total iron there is left 294 lb., equal to 378 lb. of FeO. Let us now assemble the items for the slag. They are silica 408 lb., FeO 378 lb., and lime 140 lb., constituting 90 per cent of the gl^. The resultant percentages are then computed. In the same way in the matte, since the three constituents, sulphur, iron, and copper, make 90 per cent, we can distribute the resultant percentages, i.e., 25 per cent to the sulphur, 33,5 per cent to the copper, and 31.5 per cent to the iron. The grade of the matte in copper is then 33.5 per cent.
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CALCULATION OF CHARGE CHARGB4HEET. REVEaBERATOBY SMELTING.
Dry
Co.
ao,.
F«.
c>o.
'
Lb.
cS:.
Lb.
S.
Lb.
&.
Lb.
Lb,
1Ss
e
M
S
10
«i
So
'
Cui
■Uc
s
lr„.».
-4a
Cu in lulta
Iig
Forilx- FtOIoiiUc-
IndM Wt. ol mitU
:
SH). -408 -3ft.8 jm Mot. FeO -378 -38.8 per ant. C»O-140-W,flp««M.
gr""':
per OB
M.0%
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Chapter Xxxi
Converting Copper Matte
Princ^le of the Process. — This consists in treating molten matte in a. converter, a receptacle lined with refractory material- Compreased air, blown throu^ the molten bath, as in steel converting, btims off the sulphur as S02 and oxidizes the iron to FeO, this entering the slag. The slag is pow^ off, leaving blister-copper, and this in turn is poured from the converter into moldB.
The Copper Converter
Two types are in general use, the horizontal or barrel type, and the vertical or Great-Falb type. Both are electrically operated.
The Horizontal Converter. — Fig. 217 is a view of the horizontal or barrel type, at first everywhere used. It consists of a cylindrical steel shell, having riding rings at each end by which it is carried, and revolved on four carrying rollers. By means of foxir links riveted to the converter top, the steel shell can be lifted from its stand and transferred where desired bj' 40-ton traveling crane for relining, a newly lined shell then being put in its place. To obtain better access to the interior the separate noae can be
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The Copper Converter 401
unbolted and removed. At the front is seen the rectangular wind-box, havii^E fourteen tuyeres. As better shown in Fig. 218, the air supply enters through a sleeve connection at the axis of the shell, and through a ca8t>-iron pasaf^ to the wind-box. The motor, through a worm and wonn gear, revolves the converter to any desired position. Immediately at the front in Fig. 217 is shown a band-brake by which the motor can be quickly, stopped.
From the "blast main,'' Fig. 224 as marked, a 12^. branch leads to the axial line of the converter, having there a sleeve. The air passing from this point curves around the converter to the wind box where are the tuyeres, and below them the " puncher's platform." A hinged platform at
Fio. 218— (Jreat I'alla or l.i|>rifclit I'ype Converter.
the front is used -when Epent-accretions are to be cut away or for throwing in cold matte, etc.
The Smith-Pierce is another type of horizontal converter, basic lined, and in successful use at several copper plants. They are made up to 13 ft. in "diameter by 30 ft. long. (See Fig. 220.)
The U|tti^t Converter.— This, called also the Great-Falls type, since it was there developed, is built in sizes up to 20 ft. diameter, though the 12-ft. converter is common. As shown in Figs, 218 and 219, it is cylindi ical in plan with a tapering bottom and top. The converter-top comes apart just above the trunnions, so that, when the converter has been removed for relining, this may be unbolted and lifted off. Riding-rings are bdted to the shell, and are carried on rollers as in the horizontal type. The wind-box occupies the front half of the circumference, and its connection to the blast-main, also the control valves, b plainly shown. There are
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402 "CONVERTING "COPPER-MATra:
twenty-four mdividxial tuyeres branching from the wind-box, any one of which can be aeparately removed if deared. The hole through the thung for each is 1^ to H in- In the view, Fig. 21S, can be seen the end of the worm drive, but not the electric motor.
The Converter Likihg
Both acid and haae linings have been used, the first eixiasively for many years. It was contended that the add Uning was neceaeaiy in order to furnish silica for the slag neceeaaiily produced as the iron was oxidized. An acid lining would last from acvrai to nine heat, or much leas than twenty-four houis, and then had to be removed for relining. This was done, using a ganister of 85 per cent silica and 15 per cent clay; tJie converter was then dried and heated for re-use. The lower the grade of the matte, that is the hi^^ it was in iron, the shorter the life of the lining, so that the converting of low-grade mattes was prohibitory.
It was found that by supplying silica to the charge directly upon llie surface of the molten bath the operation could be carried forward with a baac lining, ajid bo, mnce 1911, the acid linii^ has FiQ. 2I9.-Sections of IWt. BmLc Converter. ^*«» P^en up in favOT tt
the basic.
The Basic Lining. — The usual practice is to line the converter with magnei^te brick to a thickness varying from 24 in. at the
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tuyeres to 9 in. elsewhere. But the temperature in the converter is continually varying; the lining is hot during the blowing period, and cools during pouril^ or lechajgiDg. These variations in temperature cause the lining to crack and spall off, also there is the mechanical wear of the chai^, contributing to shortening its life. This led to the idea of forming a protective coating. In blowing the initial char^ of matte without silica and at a moderate temperature, it was found that above the iron oxide necessary to combine with any silica present there was an excess of iron, which under the oxidizing action of the blast, was converted to magnetite (Fe304). At about 1200" C. this magnetite becomes mushy and attaches itself to the lining, forming the needed protective coatii^, which can by control of the temperatLire be added to at will. With proper operation and care, the lining, thus protected, should last indefinitely. It has the farther advantage that low-grade mattes can be readily treated.
Fia. 220. — PieTce-Smith Horizontal Converter.
OPERATIOn OF THE BASIC CONVERTER
The operation of the horizontal converter may be thus described. The initial charge is 60 tons of matte to which is added 10 per cent of dried quartz for flxixing. The blast is now turned on for thirty to thirty-five minutes and the charge begins to heat up owing to the active oxidation of the FeS of the matte. The converter then is tilted to pour slag, the Uast being at the same time shut off. With the completion of the pouring a ladleful or 6 tons is introduced and 3 tons of silicious ore added. Another blow then begs. This cycle of blowing and addition of matte and silicious fhix is continued until 70 to 80 tons of blister-copper has accumulated. It means the charging of 300 to 400 tons of matte, and
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a period of thirty to fifty houiB of blowing time according to tiie gnde of the matte. The molten copper is then poured into lai^, hot, lined ladles and transfetiBd to the casting-furnace or dii«ctly to the c&eting-
Operatbig Pncautions. — ^As ctmverting is primarily an <mdation proceed, the speed of workii^ depends on the speed of blowing. Hence the tuyeres, which tend to slag over if imtouched, should be kept well punched by the insertion of a punching bar, as is done in blast-fumace smelting, except that in converter practice this is done every few mimttes. In this way the tuyeros aic kept open, bright, and in working order. The temperature of the converter can be r^ulated by the addition of cold matte, oie and the rich sweepings that get spilled, and that accumulate during operatitnuL An addition of a small amount of hot matte, about ten minutes before tlus charge is finished, will ensure hot copper when pouring.
Introdoction (rf Silicioas Ore to the Converter. — This has been done largely by using charging boats or trays that will hold 1 or 2 tons of material. These are provided with chains by which they are lifted and handled.
CHBHICAL REACnOnS OF THE CONVERTER
The Slagging Period. — Referrii^ to operations in a 12-ft. ccoiverter, the matte xa&y be taken as containing Cu, 43 per cent; Fe, 29 per cent, and S, 14 per cent; corresponding to ("uzS and FeS wi& a little F^Oi. The period begins when the first ladlef ul of matte has been poured in and the blast tunwd on. The heat at first drives off elemental sulphur according to the reaction:
Here is revealed the soiuve of the material to furnish a magnetite lining. But a subsidiary reaction then takes place 1:^ which this magnetite a reduced to FeO, then entering the slag:
(3) Fe304+Fe5=4FeO+4P^S. Tbe principal reaction now becomes:
In this vigorous reaction much sulphur dioxide is evolved, the magnetitB becomes reduced, according to reaction (3) into FeO, and this together with that resulting from reaction (2), also enters the sl^. It is thus seen how necessary it is at this stage to add sihcious ore to unite with tbe FeO. The reactions, at first slow, rapidly increase until in about forty-five minutes they are completed; much of the sulphur is gone and most of the iron
Chemical Reactions Of The Converter
slagged, and the converter contents brought to the stage of white metal. Tbe end of the stage is known by the appearance of the issuing flame, the greenish border, at first seen, changing to a pale permanent blue. Piecea of matte are thrown into the converter if it is wished to make the charge hotter, while sweeFongs from aroimd the converter, rich in copper, when added tend to make it cooler. At this time the converter is turned down and the slag poured into the slag-ladle. To tell when the matte begins to escape, " the skimmer " passes a rabble through the flowing stream and can thus determine the presence of drops of matte, whereupon the converter is returned to blowing podtion. This converter slag, containing 1.5 to 2 per cent copper and about 0.5 to 0. 1 os. silver per ton, is sent to the blast-furnace for recovery of its metal contents.
Conversion of White Metal to Copper. — At the beginning of this blow there is but little iron left, and the matte has been brought to the stage of white metal of 75 per cent Cu, and we have:
Cu2S+20-2Cu+SOa, 4CuaS+90 = 6Cu + Cu20 +4S0a,
FlO. 131. — Elimination of Impuritiet
n Converting.
with a further abundant evolution of SO2 and the formation of blister-copper. Tbe CuaO of reaction (6) is soon reduced to Cu as per reaction (5). As far as the agitation of the blast permits, the molten contents separate into layers, an increasing layer of white metal above, and slag on top. The Uast enters the bath horizontally 6 to 12 in. above the bottom, and blows largely through the white-metal layer. The escaping flame is white, gradually changing to rose-red and finally to a brownish red. It decreases, until at last there is but a brick-red flickering. The identification of the finish needs care and experience. If carried too far we have over-blown copper. The converter is now turned down, the blast being at the same time shut-off, and the blister is poured, either directly into molds, or into a ladle, to be taken to a tiltiog-fumace. The converter is then turned back to receiving position for treatment of a new charge. The air is supplied at a pteseure of 12 to 15 lb. per square inch, using 150,000 cu. ft. of
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air per ton of blister produced. It takes four hours for a cycle of operations, and in this time 25 tons of matte are treated, yielding 10 tons of blister-copper.
Loss in Converting.— The escaping gases contain nitro^en, sulphur
dioxide, and traces of volatiUzed metals. The loss of gold is small; that
of silver depends upon the amount of volatile metals. In the flue-dust in
one case there was an averse of 40 oz. per ton. The loss in converting
may be given at 1 to 1.5 per cent of the copper and 2 to 2.5 per cent of
the silver. When treating leady matte from a silver-lead fumaoe, this
loss of silver may be serious,
amounting to 33 to 40 per cent.
At Tooele, Utah, where the lead
bag-house, the diver is saved.
We give, in Fig. 221, a graphic chart showing the losses in an acid-lined converter where the period of the blow was seventy minutes. It shows that after ten minutes the 1.2 per cent of zinc is gradually burned off. Antimony and arsenic, both present to the extent of 0.25 per cent in one case and to 0.15 per cent in another, are well eliminated toward the end.
Crane Ladle. — Fig. 222 is a view of a steel ladle used for transferring matte or blister-cop-
^ „„ per by means of the traveling
crane from the forchearth of a
blast-furnace or from a reverberatory furnace. When in position it is
tipped by an auxiliary hoist of the crane which hooks into the eye shown
at the left side of the ladle. The ladle is plastered on the inside with a
coating of clayey loam, which is dried out and heated before using.
Blast-Furnace Smelting And Converting Plant
Fig. 224 represents the plan of a blastr-fumace plant with a two-stand converter-plant attached, the size of the latter being indicated by the number of stands or stalls in which copper matte can be blown. In the elevation. Fig. 223, the receiving track for coke is shown at the extreme right of the illustration unloading into the coke tons beneath. The
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BLAST-FURNACE AND CONVERTER BUILDtNG 407
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ore-bins with the inclined bottoms are shown to be on the same level In the furnace building there are two matting btastr-fumaces, each 42 by 144 in. at the tuyere level and each having a settler or fore-hearth 10 ft.
diameter. Blast is furnished to the furnaces from a power^ouse, not shown.
In Fig. 223 is shown the semi-elliptical flue 3 ft. by 7 ft. high, leading
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The Cottrell Electrostatic Theater 409
from the blast-funuiceB. This croeses the near end of the furnace building, as indicated by the dotted linea in Fig. 224. It is connected to a dufitchamber which leads to a stack. The slag is taken away over an electrio trolley system entering the building. The slag-cars are brought close to the settlers to receive the flowing slag while the matte, as needed, is tapped from a lower tap-hole into the steel ladle for transferring to the converters. A platfonn elevator, at the end of the furnace building, elevates slt^ and
Fig. 225.— Cottreil Tretiter.
other material from the floor of the converter building to the cfaai^e-floor for the blast-furnace.
The converter building has at one end the Uning floor, and is commanded from end to end, by a 40-ton electric traveling crane which serves to handle the converters, to supply them with matte, to take away the bU^, and to handle all materials for and from the converters. For each stand there should be an extra shell, or four in all. The escaping gases from the converters are received in a hood attached to a dust chamber
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eo that the particles of matte blown out by the blast are collected. pTom the hood a flue counecta with a dust chamber and this with a stack. At one end of the furnace building (see Fig. 224) is the mill where the lining material is prepared for the converter (now no longer used).
This, called also the Cottrell procesB, has come into use since it ie poeable by it to remove the dust and fume from the hot gases produced in roasting fine ore, or the fumes escaping at the converter-plant where, due to the h^ temperature, a b^-house could not be used.
In principle, the fume and dust in suspension flow upward throng a vertical tube 5 to 10 in. in diameter by, say, 15 ft. long, as shown in Fig. 226. No. 10 insulated copper wire, suspended axially, takes a tugh-voltt^ undiiectional current of 25,000 to 60,000 volts or more. The current, passing from the wire to the inner surface of the tube, electrifies the fume or dust particles negatively, and these are repelled to the inner surface of the tube, forming a coating upon it. This coating is occasionaUy removed by jarring it off into a hopper below, Flo, 226.^Uiim;rHiu of Trcai«r. whence it is removed by a spiral-screw conveyor. Another method, called the plate Gystem, consists in having vertical corrugated plates about 10 in. apart and at 10 in. intervals chains of J in. diameter suspended. The current down those chains acts in a similar way, repelling the dust particles to the surface of the plates. A later variation of this consists in having rods passed horizontally equidistant between the plates to carry the current. A chamber full of pipes or plates with the rods thus insulated is called a Cottrell treater unit. The insulated wires are connected in one and receive the high-pressure current.
THE WORKS OF THE nTTERHATIONAI, SMBLTIHG CO.
This comprises two separate installations, a roasting and drying plant and a reverberatory smelting and converting plant. The dried or roasted product of the first being smelted at the second, we have, therefore:
(1) A roaster or dryer plant with a Cottrell treater.
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Boasting Plant 4H
(2) A smelting and coQvertcr-pIant where the converter dust is caught in a Cottrell treater.
At the dryers or roasters (see Fig. 227) are trippers that regularly feed to five-hearth Wedge roasters. Beneath are two rows of calcine hoppers, so placed that the calcine can be drawn off into cars that take it away to the smelting works. In Fig. 236 b shown one of the two fire-boxes which heat the roasters, two and two, respectively. The roaster arms of the furnace are cooled by air and the delivery of this air is by an underground pipe marked " cooling air " in Fig. 236. Above and between the roasters is the gas flue, which receives the branch pipes from the roaster. From the top of the gas flue are pipes that branch right and left to the respective gas treaters for each roaster. E^h treater has thirty-six pipes 12 in, diameter by 15 ft. loi^ with an upward flow of gases through them. Here the dust is separated, and the gas escapes by stacks or chinmeyB, three to each treater, high above the building.
The Roaster Plant.— This Fh.-. 227— End Elevation of Roaster Plant,
roasting, or rather drying
plant of the company is for the drying of a flotation concentrate, to which is added a small amount of sulphide ore in order to produce the needed proportion of matte in the subsequent smelting. The material is so fine t^t a Cottrell installation bad to be added in order to prevent excessive loss of flue dust.
Figs. 227 and 228 are two views of the roaster building, which contains five Wedge roasters, each roaster having its own treater. In the other end of the building is the " electrical machinery room " containing the switch boards, generator-exciter sets and the motor-generator-transformer sets.
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Here are the elevating and conveying belts for delivery of the flotation concentrates to the furnaces. There is so Uttle sulphur in this product that there is no need to roast it, and the wet material is simply dried before being sent to the reverberatory for smelting. Were it needed, roasting could be added. The material being bo fine much flue dust is necessarily made.
The Smelting and Converting Plant. — Fig. 229 is a plan of this smelting plant for the treatment of a flotation concentrate, to which has been added
Fio, 228.— Side Elevation of RoBBter Plant.
a small amount of sulphide in order to produce the needed proportion of matte. The plant consists of the converter house, its Cottrell treater and stack at the right, the rcverberatory building at the center, and the boiler bouse for the waste-heat boiler with their flue and stack at the left, also auxiliary equipment of the plant above.
There are three large reverberatory furnaces (oil-fired), over which tnnaveisely run a double line of tracks which bring in the calcines from the
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Reverberatory Smelting And Converting Plant 413
diyer and roaster building, Eome distance away. The firing end of the reverberatories adjoins the converter house set at a lower level, so that the matte ladles can be set low enough to take the matte when tapped. At the front end are seen the slag cars on a sunken track. As fast as filled they are removed by a locomotive.
Waste-heat Boilers. — Over this sunken track is the heater flue, which
Fig. 229. — PlaD of Reverberatory SmeltiDg abd Converting Pl&nt.
takes the gases from the three furnaces. From this flue there are six branch flues to the six waste-heat boilers, and each of these in turn branches to the reverberatory fiue, this latter leading to the main stack 300 ft. high by 25 ft. diameter.
The Cottrell Treater for the Converters. — In the converter house are five converter stands, each arranged as in Fig. 229, the goose-neck branch pipe leading to a dust bin, the hopper of a Cottrell treater where a part
of the dust settles out. The rest of it is taken out by a converter Cottidl syatem so that the valuable dust is quite tecoveied. The ^aes pass away by the converter stack. Beneath the dust bin is a dust track and parallel to it a copper bullion track for the removal of flue dust and the blister copper respectively.
The liltiiig Furnaces. — There are two of these "See. 1 and 2 of the {dan, F^. 229, so that one furnace is filling while the other is pouring. They resemble a lai^ horisontal converter and receive a niunber of ladlesful of blister as this is made at the converters. When full, the tilting furnace is poled to make a smooth ingot and is poiued into the casting macliine adjoining, see Figs. 242 and 243. The ingots, as they fall into the water bosh of the casting machine, are there cooled, then, by an endless chain are raised and dehvered into the casting ebed for weighing and shippiug away by the copper bullion track.
Otiier Bqu^tment — ^The converter air mun brings in the air from the compressor in the power bouse (not shown) at a pressure ttf 15 lb. per square inch for use at the converters. A battery of ei^ oil tanks supplies the reverberatory fxunaces throu^ an 8^. main. Near these tanks is seen the installation of " mud bins " where clay is stored ior mixing with crushed silica ore in the silica bin. These are mixed to form ganister in a Carlin mill for use in lining the converters. They are now little iised under conditions of modem practice. The skull breaker consists of a strong grated hopper into which the skulls or shells, that form like a lining on the interior of the ladles, are broken. A wei^t lifted by the traveling crane is let fall upon these skulls, breaking them to a size for convenient haudhng so that the pieces can be charged into the furnace for melting down.
COSTS OF. A PROPOSED PLAITT AITD OPBRATIOlf
For the year 1919 we give these coats tor a proposed plant for the Consolidated Copper Mines Co. to be built at Kimberly, Nev. For this Frederick Laist, in charge oi the works at Anaconda, Mont., who planned it, prescribes a single reverberatory-fumace plant of 445 tons daily capacity, using a charge of these itemih
(1) 133 tons of concentrates of the compoation Cu, 18 per cent; SiOa, 20 per cent; Fe, 25 per cent, yielded from 2000 tons daily of " porphyry ore " of 1.2 Cu.
(2) 37 tons concentrates of the composition Cu, 10.5 per cent; SiOs, 15 per cent; Fe, 32 per cent, and S, 34 per cent, made from ISO tons " sulphide " ore contabing 2.9 copper,
(3) 150 tons of oxidized ore containing 7.5 per cent copper,
(4) 125 tons of fluxing materials.
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Estimated Cost Of Reduction Works 415
It is assumed that these ores can be delivoied to the reduction vrorks at a cost of $1.10 iox the porphyry ore, S5 for the sulphide ore, and tlO for the oxidtzed ore.
The reduction worits needed for treating this quantity of ore consists of a concentrating plant having a capacity for 2000 tons of porphyry ore and 150 taoB of sulphide ore also a power [dant capable of generatii^ 3000 kw. which will furnish the power needed for water supply, operating the minii^, concentrating and smelting plants.
The cost of the water supply is h^h, mnce it must be pumped to a total bei^t of 967 ft., or against a total head of 1200 ft. and a distance of 13 miles.
The coDcentrator is to be constructed m two sections and provided with E^re grinding mills and flotation machines, so that a breakdown of one of these will not affect operations.
The smelting plant mill comprises four 20 ft., Beven^-hearth Wedge roasters; one 100 ft. by 20 ft. reverberatory furnace equipped with waste heat boilers and a converting department containing two 12^t. upright converters. The smelter buildii^ should be made of size to accommodate . another reverberatory furnace. Such extension would cost $50,000, while a second furnace would cost $125,000 more.
Sstimatbd Cost Of Kbduction Woiucs As Odtiiited Herewith
AaBume daily treatment™ Porphyry ore 2,000 tana
" Sulphidct " ore 150 toiu
" Oxidiied " ore IfiO Uaia
Water supply 1000 to 1500 gal. per minute $305,740
Cniflhing plant 100,000
ConoentTstJng plant 450,000
Power pUntB— 3000 kw 460,000
Drying plant 200,000
Reverberatory plant 350,000
Converting plant 150,000
Bins, rolling stock, ebope, houses and miscellaneous 300,000
S2,30S,740 Engineering, drafting and contingency (say) 194,200
Total $2,600,000
This estimate is based on present cost of supphes and labor and is considered conservative.
The plant can be enlai^ed at any time, without interfering with operations, to 5000 tons of porphyry and 250 tons of oxidised ore, at an additional expense of about $1,400,000.
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OPERATING EXPENSES Assume smelter recovery at 95 jier cent, uper&ting coeto are estimated m foDows. Power — SO.Ol per kw.-hr. ConcentratinK— 10,85 per ton. Drying — 10.45 per ton, Reverberatory Bmelting — J2,35 per ton. Converting and casting — $10 per ton of <upper.
Crushing 276 tons ore, flux and »e<'ondari(« at IS eentB,r $41.20
Rofistmg445 tons at tO,4S 200,00
Reverberatory smcltinR 445 tons at 12,35 1,045,00
Converting 37 tona Cu at 17,50 278 00
Casting and loading 27 lone at S2.S0 92.60
$1,656,80 Add 10 per cent for miscellaneous 165 68
S1,S22 48
Cost of smelting per Ion ore concentrate mixture 6 70
Coat of smelting per lb. of copper produced 0.0246
Coat of concentrating per lb, of copper produced 0 , 0247
Treatment of Porphyry Ores Alone. — In order to convey an idea of the value of the porphyry ores alone, without admixture of high-grade " oxidized " ore, the following estimate is Bubmitled. The complete treatment of porphyry ore alone locally would scarcely be feasible on a scale of much less than 5000 tons per day. This quantity of ore would, however, yield approximately 300 tons of concentrateB, which would make the total amount of material to be smelted, including fluxes and secondaries, about 425 tons per day, which would be an economical operation for one reverberatory furnace.
The cost of the operation would be about as follows:
Assume recovery in bullion of 80 per cent of 1.4 per cent Cu=22.4 Ib. per ton.
Cost Of Tkeathent Porphyry Ore Alohe— 6000 Tons Per Day
Pm Lb. Co
Mining at SI .10 per ton of ore SO . 0491
Concentrating at S0,S5 per ton of ore . 0379
Smelting at S6 per ton of concentrates 0179
Freight at SI6.60 per Ion copper 0083
Refining at S22 per ton copper 0110
Gross cost, , ■, 10.1242
Credit for gold and silver .0050
Net cost SO. 1192
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Chapter Xxxii
The Hydrohetallurgt Of Copper
PSmCIPLBS OF THE HYDROUETAU.UHOY OF COPPER
The wet methods of extractii^ copjier from cupiferous ore consist in obtaining the copper from the crushed and perhaps roast«d ore, in water solution, either with or without the aid of other solvents such as a solution of ferric oxide or of sulphuric or hydrochloric acids. The copper must be in combination with elements that will permit it to dissolve in the solvents used. Thus, metallic copper would not dissolve in sulphuric acid, and chiysacolla is difBcultly soluble. From the clear decanted or filtered copper-bearing solution the mRtal may be precipitated electrolytically, or with scrap-iron or with lime. The resultant " precipitate " is then melted and refined.
Avsilable Copper Ores. — Copper has been extracted profitably from suitable ore of as low grade as 0.5 to 1.5 per cent copper when the conditions of an abundant and easily exploited supply and cheap labor prevailed. Such bodies of ore, much of it in oxidized form, occur throughout the world, often of too low grade to be treated by smeltii^, or too difficult of access. Besides this there are huge dumps, being the tatliags of concentratii^ mills. Ore containing the copper as oxide, carbonate or sulphate is best suited to extraction, but if containing lime, magnesia, ferrous oxide, or manganese oxide, it is less desirable. Copper-bearing sulphides may be profitably treated for the extraction of the metal, the ore being oxidized by weathering until the sulphide has been changed into a sulphate soluble in water. Sulphide ores may be roasted with salt to bring them into form of a chloride, which then is extracted with brine solution.
Advantages of Leacliing.~It would seem that for low^rade ores leaching should be superior to other methods, since the worthless gangue, which is the largest constituent of the ore, remains untouched, and the solvent acts on the relatively small quantity of valuable metal. Particularly does thia seem to be true of silicious ores, the silica of which in no way interferes with the leaching, while they are expensive to smelt.
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THE HYDRORlETALLUIiaY OP COPPER
BXTRACnoH OP COPPER BY HATDKAL OR WBATHBRIKO UETHODS
(1) By direct treatment of the raw or crude ore (Rio Tinto procefls).
(2) By treatment of the ore which has been subjected to a prdiminaiy roast in heaps (Shonntm Copper Co. proceas).
Outline <d Process. — ^The copper in the sulphide ore is brou^ into soluble fonn as sulphate, and frc»n the filtrate the copper is piodiHtated by means of scrap iron. The copper-bearing pyrite is made into la^;e fiat-topped heaps which are oxidized by means of a regulated supqily of water and air, and when the copper has been changed into sulphate, the material is leached with water to extract this. The clear scdution is conducted to tanks filled with p^ iron wherB the copper is precipitated.
When the copper in the ore occurs as chalcopyrite (CuFeS2) or as cord-lite (CuS) oxidation taY>ceeds slowly and imperfectly uid, for successful working, it should be in the form of chalcocite or copper ^ance (CusS). It is because of the extent of the ore bodies and the cheapness of labor that the Rio Tinto process has been successful. This ore contains on an av^age 2 per cent copper.
Preparation of Site and Heaps. — ^A site is chosen upon impervious sloiung ground for suitably draining off the sohition as formed. A clay^ or rocky bottom is required, or one properly puddled or coated with diQr to render it impermeable. The heaps may contain 100,000 tons of ore and are constructed as follows: On the ground is first arranged a network ctf flues 12 in. square, made of lump ore. Vertical flee or chimn^s tiiat connect with the ground flues are built 50 ft. apart as the heap is made. The ore is broken to 3 or 4 in. diameter and some of the limipa are screened out for making the flues, leaving some fine. The run of ore is dumped and spread on the site in layers until this is 30 ft. hi^. The flat-top surface, having a grade of one in 300, is formed into 20 ft. squares by ridges of fine ore so as to ensure distribution of water within specified limits, and wadi-ing of the pile from the top to the ground flee or drains. Launders are provided to carry water to the heap.
First Operation. — As the heap is forming, water is applied to extract any already-formed copper sulphate. Oxidation starts as the result of the wetting. The completed and wetted heap begins to oxidise r^dly, aa shown by the heat evolved, the temperature of the air in the chimneys risii^ to 70° C. As the heat increases the ground flues are closed to control oxidation, and to spread the reactions through the heap. The surface assumes a brown color, due to the dehydration of the baac fejric salt that forms, and heating is made apparent by this drying action. Great care is taken to prevent the heap frran catching fire.
Rio Tinto Process 419
Chemistry of the Process. — ^By the combined action of sir and moisture the following reactions occur:
that is, pyrite ie oxidized to ferrous sulphate and sulphuric acid. This ferrous sulphate is readily oxidized to ferric sulphate thus:
The tunis-formed ferric sulphate acts on chalcocite and changes it in part to copper sulphate, itself reverting to ferrous sulphate according to this reaction:
The cupric sulphide, hitherto unaffected is farther changed as follows;
(4) Fez(S04)3+CuS+30+Ha0-=CuSO4+2FeS0*+H»S04.
Reaction (0) is relatively rapid, and accordingly about half the copper goes into solution in a few months. Reaction (10) is slow, but in two years, under fovorable conditions, yields 80 per cent of the remaining half of the copper.
Extraction. — ^When oxidation has advanced as far as is safe, water is applied at the rate of 220 gaL per minute until the soluble copper salts aie extracted. The flow is then stopped and oxidation is resumed, and is foUowed by renewed washings. After a year the top-surface needs retilling; the ridges are arranged where the squares formerly were and the launders are shifted to conform. At the sides of the heap for the distance of some yards the ore has become cemented, and holds copper salts. These aides arc dug down in terraces to expose the copper salts and to extract them by washing. When there remains but 0.3 per cent copper, extraction is considered complete. This pyrite heap, after the copper has been washed out, is still valuaUe as a sulphur-bearing l^te, and many tons of such washed ore has been shipped away to the sulphuric acid makers.'
Reduction (d Ferric Su^lhftte. — ^The solution that flows from the heap contains ferric sulphate, and to prevent it from consuming iron in the precipitation tanks it must be reduced by running the Uquor through a filtei^ bed of fre^ iion-pyrite smalls or fines. The reaction is as follows:
(5) 71^ (S04)3 + FeSa + H3SO4 = 1 5FeS04 + SHaSO*.
This filter-bed is retained within a reservoir formed by a masonry dam across a small ravine. The hquor, or solution after percolating the bed, flows to a common settling tank. It is then drawn off to a series of tanks, canals, or chimes containing pig inm. The typical solution entering the series would oontun CuO, 4 per cent; FeaOs, 0.1 per cent; FeO, 2.0 per -
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420 The Hydrometallurgy Of Copper
cent; HzSO*, 1.0 per cent and Ab, 0.03 per cent. The presence of so much FeO and H2SO1 ia due to the fact that a part of the waste or barren aolution, leaving the eeries, is pumped back and used for wateiing the heaps, so that the sohition tends to increase in these elements.
Precipitation. — ^The copper-bearing liquor or solution, drawn bwa the filter-bed, is run through precipitation launders coi)taining {Hg4ron ingots piled in open order, and the copper is precipitated (replacing the irm which dissolves) in the form of " cement copper " or copper precipitate. Following the tanks the solution enters the canals, flumes or launders, arranged on the slope of a hill in such fashion that the solution may pass back and forth through them until it is discharged " barren " or free from cc^per from the end of the lowest series. These flumes are 320 ft. loi^, Sj ft. vHb by 2J ft. deep. They begin at a grade of 2 in 1000, and the final (himes slope 11 in 1000. The rate of flow is thus increased and less pig iron is wasted. Some of the flumes are cut out from the flow or by-passed daily, the solution meanwhile going through the remaining ones. Those thus cut out are drained, and all the pig iron is removed and piled beside them, the copper attached to them being meanwhile knocked off and thrown back. The muddy precipitate at the bottom of the by-pasaed tanks snd flumes is removed to the cleaning and concentrating plant, while the [ng iron is piled back, and the flow of solution again directed through the flumes. Under the best conditions there is needed 1.4 tons of pig inm per ton of copper precipitated.
The first reaction m the tanks is that between imreduced fenic sulphate and the pig iron, thus expressed:
'i
It ia a reaction that wastes iron. Precipitation of the copper is iHOU^t about by an electro-chemical reaction, viz.,
Finally a reaction takes place between the free sulphuric acid and the iron.
This reaction is evinced by bubbles of hydrogen rising thiouf^ the tank liquor. It means a further waste of iron.
Treatment of ttie Precipitate. — At the cleaning plant, the crude precipitate, containing 70 per cent copper, by means of a strong jet of water, is gradually worked over and through a coppei^plate screen, this screen being situated at the bead of a long launder. The overmze of the screen,
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Shannon Copper C30Mpany Process 421
consistmg of le&f copper and smaU pieces of iron, ia thrown into a heap to be picked over by girls who remove the scrap iron. The fine passes through the copper-plate screen, and is turned over by a stream of water that washes out the dirt and light particles, leaving the copper behind.
At the head of the wa^iing launder for a few yards is found No. 1 — precipitate of 94 per ceat Cu and 0.3 per cent As. Farther along is No. 2 — preci|ntate of 92 per cent Cu. Next comes No. 3 — precipitate, which is fine, and contains 50 per cent Cu, 6 per cent As, some graphite (from the pig iron), and the bismuth and antimony precipitated from the hquor. Noe. 1 and 2 precipitate are sacked for shipment, and No. 3 is added to a blast-furnace matting charge, the copper combining to form matte, while the impurities mostly volatilize.
It has been urged against the FiG.221,-Method of Removing the
Rio Tinto process that it is a rather Cement Copper,
complicated and very lengthy process, and that it ties up too much capital. However, where labor is cheap, and forms the principal expense, where the ore is suitable and abundant, where the climatic conditions are favorable and water-supply sufficient, this, as experience has shown, seems to be the most practical of the methods thus far evolved. The process has been in use for centuries.
(2) THE SHAHIfON COI9BR CO. PROCESS
This process is principally used in treating oxidized ore together with sulphide. A typical oxidized ore may contain 1.9 per cent Cu, 40.8 per cent SiOs, 16.5 per cent Fe; and 15.4 per cent of the alkali earths. If subjected to sulphuric acid leaching it would need 7.5 to 8.8 lb. of acid per pound of copper recovered and this proved so expensive that the present process was devised.
In operation, 1000 tons of the oxidized ore, crushed to 2-in. size, is piled on 100 tons of sulphide ore in circular heaps, and ore fines are added as a cover to the thickness of 1 ft. There are ground flues and stove pipes placed vertically, reaching from the sulphide bottom to the top of the heaps, all to provide draft for the burning of the pile. The sulphide ore, having been set on fire, evolves SO2 and SO3 gases which rise through the entire pile. At the same lime the barren hquor from the precipitation tanks is sprinkled upon the heap. Reactions take place between the rising gases.
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422 THE HYDROMETALLURGY OF OOPraSI
the iron sulphate ttolution and the oxides and carbtoiatea of tiie oie, teplesented by the followii^:
That is to eay, ferric sulphate is alternately reduced to ferrous form by SOj and the resultant ferrous salt again oxidized. In the reduction sulphunc acid is set free, and this is ready to dissolve the bases and the copper oxide. The sulphur trioxide combines with the bases directly, forming sulphates. Toward the end of the heap-treatment the pile contains much ferric sulphate, an effective solvent for basic sulphates and unaltered carbonates. Wbere the roast gases most effectively penetrate the pile, some 85 to 96 per cent ftf ihe copper is brought into soluble fonn, but the considerable quantity of clayey material in the ore causes clogging and hence imperfect action in sections of ihe heap.
The ore is transferred to circular tanks 25 ft diameter, 5 ft. hi^ and holding 75 tons each. These have a filter^bottom covered with cocoa matting. The ore, coarsely crushed, as already stated, is early leached with water by percolation.
The copper sohition from the leaching tanks, now containing much ferric sulphate, is run through a bed of raw oxidized ore, followed by a bed of sulphide ore or of tailings containing sulphide. The action of the ferric sulphate on the oxidized ore is to dissolve its contained copper, while the excess of ferric salt is later reduced to ferrous form as the result of its ctm-tact with the sulphides.
The copper liquors, as in the Bio Tinto process, are run into flumes or launders 300 ft. loi^, 5 ft. wide, and 2 ft. deep, contuning acrap iron. The spent liquors &(Hn the launders carry as much as 3.5 per cent Fe as ferric and ferrous sulphate. The extraction raises from 73 to 82 per cent of the contained copper. The copper precipitate is removed from the launders in the same way as in the Bio Tinto process.
BXTRACnON OF COPPER AS A CHLOSIDB
By this method the ore, after crushing to 4-mesh size, is given a roast by which -psri ot the sulphur is expelled. At this stage salt is added and the ore is finished by a chloridizing roast. After cooling, the roasted material is leached with a salt solution to extract the copper as chloride. The copper in the filtrate is precipitated on scrap iron. The process has the merit, that where gold and silver are present, they may also be dissolved and recovered. Ore of as h^ as 70 per cent siUca, containii^ sufiSdent pyrite or chalcopyrite, and crushed to l&^nesh, can be given a
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9Ender80N Process 423
fni<!ce88ful ckloridizing roast with salt. If it contains no pytite this can be added and a chloridizing roast given in from six to twelve hours, and the copper, gold,' and silver leached out by means of water and dilute acid, tha latter being obtained as a by-product of the roast.
Wc describe two methods, viz.: (1) The Henderson process, and (2) the Laist process.
1. The Hendenoo Procesa. — The weU-4t)asted residue, or .cinder, resulting from the pyrite used in tnakii^ sulphuric acid, contains 2 to 4 per cent copper, with silver and gold. All these metals can be extracted by a cbloridizing roast followed by leaching with weak liquor from a previous operation, containii^ water and dilute hydroctiloric acid. The copper in the clear filtrate is precipitated upon scrap iron.
The Plant. — Fig. 232 shows the plan and a transverse sectional elevation of a plant of the Pennsylvania Salt Manufacturing Co., Natrona, Fa.
Fi<i. 232. — Elevation of Henderson Procewi Plant.
The cinder (red-roasted or burned pyrite) that is brought from the various sulphuric-acid plants throughout the coimtry is ground dry to 2(>-mesh in a pan-mill, mixed during the grinding with 12 per cent of the weight of salt. This is raised by belt elevator to storage bins on tiie floor K and put into the roaster feed hopper S.
The mixture is sent to the five-hearth Wedge chloriziding roaster, Fig. 233. Each hearth is perforated with flues from side to side of the furnace. The flame from a firebox on one side, passing throu^ these flues, reaches the vertical flue on the opposite side which leads away to the chimney. Dampers regulate the direction of the drafts. Thus the combustion gases are kept separate from the chlorine and the acid gases generated from the ore. Raw pyrite is charged with the mixture to have the ratio of copper to sulphur as one to Ij. The hearth is maint^ned at a just viable red {525° C.) by the use of 10 per cent of fuel. The gas, from the ore being chloridized, is down-drafted from the upper to the lower hearth,
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424 The Hydrometallurgy Of Copper
then passes to Ute scrubber or condensing tower a filled with lump coke wet with a water spray. The water, in contact with the ascending gas,' abeorbe the chloride and sulphurous acid, reacting as follows:
That is, sulphuric acid and chlorine are formed, and these react in aqueous solution.
The charge when finished will contain 80 per cent of ita copper is soluble form. It is drawn out upon the floor, allowed to cool, shoveled
Pio, 233— Wedge Chloridizing Purnftoe.
into charge^are, raised by platform elevator to the charge-floor level, and put into the leaching-tanks d, each of which is 12 by 14 ft. in size.
The ore is first lixiviated with a weak Uquor from a previous operation to remove most of the copper. The solution becomes a strong eolution. The ore is then treated with water, to remove the remaining copper, and the solution becomes the weak solution of the succeeding operation. Finally, the weak solution of hydrochloric acid from the towera a is applied, dissolvii^ the cupric oxide and cuprous chloride, hitherto insoluble. The residue, called " purple ore," is shoveled from the vats to the floor c and thence discharged into the railroad cars below.
The weak solution is sent to the lixiviation tanks. The strong solution, when the specific gravity reaches 18° B., is drawn to tanks 12 by 12 by 6 ft
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436 The Hydrometalluhgy Of Copper
filled with scrap iron, where the copper ia precipitated. The taaka han false botUnuA of slats 2 ft. above the bottom. Live steam, directed into the solution, agitates it. The copper precipitating upon the inm, works down between the slats to the bottom of the tanks and is removed to tanks g, 10 by 10 by 5 ft. The solution from this tank is drawn into launders ctotaining scrap iron as a guard, and to retain any remaining particles oi precipitate. The pteciiHtate is 90 per cent copper, 35 02. alver, and 0.15 oz. gold per ton. It is sold to the Uue vitriol makers, who pay 95 per cent of the.Bilver and the full value of the copper and gold.
The coBt of treatment by the process, with common labor at U.50 per day, is 11.87 per ton of cinder treated.
2. The Laist Process. — This process is used in the treatment of mill tailings containing 0.55 Cu and 0.5 01s. Ag per ton.
The proceed consiste in giving these tailii^ an oxide^ihloride roast, using 1 per cent ctf comnum salt. The roasted ore is then leached, fint with a No. 1 or weak solution containing 3.5 per cent H3SO4 and 10 per cent salt, then with a No. 2 or strong solution of 6 per cent HgSO* and 10 per cent salt. The copper, dissolved by the weak solution, goes to taakB for predlHtatioii on scrap iron; the strong solution, after use, is returned to be employed as No. 1 or weak solution on the next charge.
In Figs. 234 and 235 are given views of the leaching plant for the treatment of 60 tons of tailing daily.
Three bins, one A for salt, one B, for coal, and a third C, having a hopper bottom, are for the storage of the mixture of sand and slime tailing (4 of sand to 1 of slime) which is to be treated. These tailings contain 0.6 per cent Cu, 82.2 per cent SiOa, 1.9 per cent Fe, 2.2. per cent S and cany 0.55 oz. Ag and 0.002 02. Au per ton.
From the sand bin the material is delivered by a belt-feeder to a vwtical elevator discharging into the feed-hopper of the 20^. MacDougall roasting furnace D. Referring to Fig. 236, the furnace is of the six-hearth type with a lower water-jacketed one for cooling the calcine or roasted ote. There are two fireboxes with shaking grates, discharging into hearth No. 2, which thus becomes a combustion chamber, where the coal gases bum with along flame and the products of combustion are drawn off by a No. U Buffalo blower to the chimney. On the upper three hearths or Boots the tailings are roasted and brought to the temperature of 540° C. falling tboi to the floor. No. 4 where 1 per cent of common salt is fed in. During tb^ pass^e over the fourth, fifth, and sixth floors the copper, as well as the silver compounds, are chloridized, the heat still present in the ore being sufficient to ensure the chloiidizing reactions. A small vohuoe of air is' drawn through hearths Nos. 6 and 7 by a No. W Bxiffalo blower to an absorption tower, £, so as te catch any copper or silver that has been vtiiar tihzed. This tower, filled with coke, is showered with water delivered to
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428 The Hydrometallurgy Of Copper
it by a H-ia. bronze pump. To withstand the chlorine fumes evolved
from the cooling ore on the seventh hearth, copper r&bbles are provided.
The cooled ore is delivered by a horizontal screw-conveyor to a vei^
Fia. 23S.— Seclinnal Elevation of Roasting Furnace.
tical elevator which discharges it upon a 12-in, belt-conveyor to the sand distributor of one of the 32 by 12 ft. " leaching tanks." The distributor, revolving about the axis of the tank, delivers the calcine e\'enly over its whole area. When ore has been roasted and its colloids destroyod, it
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is in excellent condition for leaching. The leaching tanks will hold 360 to 400 tons.
There are two lead-lined eoiution-tanks each 27 ft. diameter by 12 ft. deep. " No. 1 solution tank " contains the No. 1 or weak solution of 3} per cent SO* and 10 per cent of common salt; " No. 2 solutitm tank " carries the strong solution of 6 per cent HaSOt and 10 per cent salt. For a 320-ton charge, the No. 1 solution is run on the tailing first to get out tiie bulk of the copper, and remains in contact with it for fourteen hours. This solution, the only one precipitated, is run out by a launder to one of the 6-in. Pohle air-lifts P, P, which raises it so that it is carried to the 27 by 12-ft. " copper solution tank." From this tank it flows in a regulated stream to one of the " lead-lined precipitating laimders " at the end of the building filled with scrap iron, where the copper and silver are quite pre-ci[Htated. The spent or barren liquor flows to waste. The precipitate, when a clean-up is made, is washed down into the " clean-up tank," and the precipitate collected for further treatment.
To return to either leaching tank: Twenty-four tfms of strong rar No. 2 solution is rrm on, and stands for seventy-two hours, after which It is returned to " No. 1 solution tank " as weak or No. 1 solution. Following this the remaining values are removed witli water^washes. The final tailings are sluiced out of the tank with a 2-in. hoee, and sent to waste by means of a launder carrying a 3-in. stream d water. The arrangement of the launders and of the four discharge valves of a tank are well shown on the plan.
The percent^e of recoverable copper is 85.4 and of alver 91,1.
SULPHORIC ACID LEACHmO
This method of extraction is suited to a limited range of ores, those which will not consume much acid (generaUy lost in uselessly dissolving bases) and those in which the copper minerals are present in soluble fonn. As regards the first objection, there are oxidized copper ores containing iron oxide and especially the alicali earths which consume acid. The copper minerals unattacked by sulphuric acid are metallic copper, cuprite, fresh unaltered sulphides such as chalcopyrite and covellite, and massive chrysacoUa. When, however, the ore contuns malachite, azurite, copper oxide, and basic sulpbatee mainly, then it may be quite suited to sulphuric-acid treatment.
Although we have a suitable ore, still the acid will act on the clayey minerals, the iron oxides and the alkaline earths so that these bases should be present in ^nall quantity only. The waste of acid is not the'only drawback; the bases named accumulate in the solution, which naturally is to be re-used, and finally the whole has to be run to waste and fresh acid used.
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430 THE HYDROMETALLtJRGY OP COPHER
When, to save scrap inm, electrolytic precipitation is employed copper is indeed precifHtated, but the SOs, nleased from the electrolyte, ondina the ferrous oxide present to ferric form and this at once proceeds to redisBolve the copper. True, porous diapbn^ms have been uaed to confine the ferric oxide to ihe anodes, but this, aside from added expense, indcaaes the resutances. Farther, no anode is altogether satisfactory. If of lead, this is gradually changed to peroxide by the oxidizing effect mentioned, and, while the peroxide may be recovered and again reduced to lead, this increases wen-king costs.
The electjolyte becomes foul, owing to the accumulation of sulphates of iron and other metab, and it is necessary periodically to send some of the Bohition to waste, thus causing a loss of acid. R^eneration of the add is effected, and in fact, when sulphur dioxide is injected into the electrolyte as a depol&riser, an excess of acid is obtained by the combination ot the SOa with the nascent oxygen liberated at the anode; stJU so many leactioDB occur among the foreign metals present in the electrolyte, and so much trouble has been found in property regulating the current denaty, that a great deal of current is wasted in excess of that theoretically needed. In the depoEdtion fnaa a copper sulphate solution this would be 2.14 lb. of copper per Idlowatt'-hour, but in practice but SO per cent of this has been obtained. With a pure electrolyte and taking proper inecautions to pie-vent wastage at the cathode an efficiency of 90 per cent or 2 lb. Cu per kilowatt-hour should be attained. These precautions would consiBt in precii»tating the interferii^ metals by chemical means before the solution goes to the electrolytic cells. With such efficiency the cost, especially where hydro-electrolytic power is available, should not exceed 1 cent per pound of copper deposited. The advantages of electrolytic precipitation are that the acid is regenerated and that a pure copper is produced.
We describe herewith two methods which have worked well upon suitable ores:
The ore exists as a large body of decomposed granite carrying 2 per cent of copper as copper carbonates chrysacolla and cuprite.
Briefly stated : The crushed ore is leached with sulphuric acid, the filtrate heated to 60° C, the copper electrolytically deposited, and the remaining solution, still containing copper, strengthened with add and returned for re-use. The oie is farther leached with water, the water run through other ore, then over scrap iron to obtain the remaining copper.
The ore, crushed to suitable size for leaching as described under crude-
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Butte-Duluth Process
ing, ia fed to three rectangular leaching tanks, 70 ft. long, 12 ft. wide by 6 ft. deep, lined with sheet-lead and having plugged bottom-openings 12 in. diameter. There is a false or filter bottom made of 2-4n. planks, bored full of f-in. holes, for the passage of the solution. The leaching tanks, when filled, are treated with a 10 per cent solution of sulphuric acid, which lemains upon the ore for twenty-four hours, dissolving most of the copper. This copper-bearing solution, still containing 5 to 8 per cent H3SO4, and having 2 per cent Cu, is drawn from the tanks and passes to a storage sump. From the storage sump it is lifted by steam-lifts into the
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Fio. 237 .—Solution Flow-eheet, Butte-Dututh Mill.
temperature cells, these heating the solution to the desired temperature of 60° C. The temperature cells are lead-lined wooden tanks.
From the temperature cells the solution Soira through the electrolytic cells, where part of the copper is deposited and acid regenerated. After flowing through these cells, the H2SO4 has been raised to about 1 per cent in strength. The cells (shown as a rectangle divided into four) consist of twelve cells each 8 ft. long 30 in. wide and 39 in. deep, lined with 4-lb. hard lead. In each cell are twenty anodes of hard lead weighing 10 lb. per square foot and nineteen cathodes, being starting sheets weighing 2 to 3 lb. when first placed. These remain in the cells seven to ten days. They grow to 40 to 60 lb. briore ponoval and assay 99.96 per cent Cu. The anodes and cathodes are electrically connected in multiple, the cells in series.
432 The Hydrometallurgy Op Copper
The solution passes to the sump tanks where it is stTengthened to 10 per cent H3SO1, the acid being received from a tank marked " sulphuric acid for standardization." From the sump tanks the strengthened aohition is pumped to the storage tank pbwied 40 ft. higher, so as to connect the " leaching tanks " by gravity. The pumping is effected by the aid of two 4-in. lead-hned centrifugal pumps.
Returning to the leaching tanks: The ore has just been leached with the 10 per cent solution coming from the storage tank above. This is followed by several water-washes from the " water tank." The first Wa^-water containing sulphuric acid is added to the mill solution of the storage sump. The remaining washes, weak in acid, go to the leaching tank shown at the right side of the flow-sheet. In order to keep the mill solution puic, and prevent it from accumulating, a quantity of it, equal to the first wasb-water, is passed by the line marked " a small part sol." through ore in the ri^t- huid leaching tank until its contained acid is used up by acting on the oro, then run to the launder marked " scrap-iron precipitation "; the spent solution is wasted. Thus part of the copper is recovered electrolytically and a small part by means of scrap iron.
The costs are thus roughly given per pound of electrolytic copper produced on the treatment of 50 tons, recovering 2000 lb. of copper daily:
3J lb. acid at $27 per ton $0,04725
Power for crusKing and electrolytic deposition 0.01
Management and labor $160 daily.. O.OS
The cost for acid should be greatly reduced, and the labor costs are excesnve, due to construction work and alterations in addition to actual operating.
The ProcesB. — This, in brief, consists in leaching the ore (crushed to i-in. size) for eight days by a counter-current system; reducing the ferric iron in the resultant solution to ferrous form, using sulphurous acid gas (SO3) to do so; and electrolytically precipitating out part of the copper, , which is then returned to the leaching solution.
Coarse Crushing. — This is done during two eight-hour shifts, once there is no storage between the plant and the mine. The ore, some of 3 to 4 ft. in minimum dimension, is crushed in two sets of gyratory ore-breakers to 4-in. fflze, then deUvered to a storage-bin of 10,000 tons aggregate capacity, as is clearly shown by the general flow-sbeet. Fig. 238.
Fine Crushii^;. — ^This is done between 3 p.m. and 7 a.u., but can be kept up for twenty-four hours if necessary. The oro is crushed in two
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The Ajo Process
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434 The Hydrometallurgy Of Copper
stages by means of Symond'a disk crushers, the ore being drawn off and crushed accordii^ to the immediate needs oi the leaching plant.
Leaching. — Of the twelve tanks eleven are for leaching, the twelfth is a solution settler. Of the eleven, seven always contain ore in prooeas of leaching. Referring to the flow-wheel, we may assume that the ore in tank No. 10 is the oldest, and No. 5 the newest in the circuit; the No. 6 is being chained with ote. No. 7 is empty, No. 8 is being excavated and No. 9 in various stages of washing and draining. When tank No. 6 has been charged, and it is ready for the leaching cycle, the " add advance," that is, the amount cS acid-bearing solution that proceeds from tank to tank is increased to its rpft"""""" amount of 2000' gal. per minute for four hours, this solution being gotten from storage tank Aor E. Meanwhile the usual advance of 1000 gal. per minute continues to go from tank No. 5 to six reducing towers marked in purs Si to St, where it is subject to the reducing action of SOs in water solution. The excess of 1000 gallons is advanced into tank No. 6 until the ore is covered with it — this, in order to prevent any interruption of flow to the toweis. When the ore is covered the excess advance is cut off to the normal of 1000 gal. per hour. Sohition on the new charge is now circulated on itself, imtil it is clarified, or for about four hours. Tank No. 6 is now put in circuit and liie neutrai advance (add free solution) to the tower comes off from tank No. 6 in place of tank No. 5.
The leaching of the ore in tank No. 6, now begun, continues for seven days, during which the free acid in the solution increases from 0.5 per cent to 3.0 per cent on the seventh day. At the end of the seventh day, the " acid advance " from the tank house is transferred from tank No. 10 to tank No. 11. Upon the entrance of a new chaige into the circuit, the sohition remaining in the oldest tank is drained to solution storage, where it is standardized by additions of sulphuric acid and is later used as " acid advance." After dndning the tank is ready for wash-water. As the copper that is taken away in the leaching is about two-thirds the total, the question of thorough washing, to remove the rest, is important. Four successive washii^ with the dratnings between are used. During tbe three-hour circulation that each wash is ffvea, an equilibrium between tbe dissolved copper in the tailings and that of the woah-water being apphed, is expected to be reached. To follow more readily the method t^ washing a charge the Sow-sheet must be referred to. When tank No. 9 has been thoroughly drained, the charge is covered with wash-water from W, circiilated, then drained to solution storage tank AorE; this constitutes the first wash. H is now covered with wash-water from Wa, aimilai^ circulated and droned toW. In the same way the wash-water from Ws ie put on, circulated and drained. The fourth or last wash, consisting entirely of fresh water, is pumped, circulated and drained into wash-water tank Wi. In this manner the fourth wash of any one charge is used as
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The Ajo Process 435
the third wash of the succeeding charge, the third aa the Becond, and the second wash aa the first. In other words each wash-water is used four times, the copper contents increasing each time, when it enters the system and makes up for the continuous losses of solution due to evaporation, to discard and to about 1 1 per cent of solution taken away in the tailings. Thus, before the first wash the solution contained 2.56 per cent acid and 2.4 per cent copper, while at the end of the fourth wash there remained O.IO per cent acid and 0.38 copper. To obtain an even better extraction of the copper it has been proposed to give a fifth wash, then allowing the resultant BohitioQ to flow over scrap iron.
Amngement of die Leaching System. — The twelve leaching tanks are arranged in two rows, as shown on the flow-sheet, and more particularly as seen in the sectional tnm?vcrse section, Fig. 239. The aisle between the two rows of tanks is 108 ft. wide and contains what is called the " central structure " of the same length as the row of tanks. This consists of six heavy concrete pieis each of four pillars supporting steel trusses from .
Flo. 239. — Croes-Bection of Leaching-tanlcs.
pier to pier. The structure has two decks, the upper carrying the belt-conveyor, the lower the solution launders and the pipe-lines. At each concrete pier are four pumps and pipe connections. Underneath the central structure and parallel to it are to be seen the two drainage laimders used in carrying the sohitions from the leaching tanks to the solution storage.
The ore, from stor^e as finally crushed, is conveyed through an automatic sampling plant, and thence by a main conveying-belt the full length of the central structure. To fill any desired tank there is a tripper on this main conveyor that delivers to the belt of a Robins spreading bric^. This bridge set over any desired tank delivers its load into it from side to side and can be moved to fill any part of the tank.
Removal of Tailings. — ^After a char^ has been washed end drained, the tailings are removed by a Hulett excavator, similar to those used in unloading iron ore at the lower lake ports. A heavy steel bridge on tracks spans the leaching tanks and can travel their entire length. On this bridge tonveis the excavator, consisting of a walking-beam, bucket-leg, and bucket
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436 The Hydrometallurgy Of Copper
of 12 tona capacity. This will unload the tMik at the rate of 500 tons per hour. Two eight-car trains are released from mine-service at 4 f.u. for the transportation of tilings, and twenty-one to twenty-three train loads are required when removing the contents of a tank. When it is desind to exchange places for the spreading and excavator brii^es it is thus performed: Just beyond the last tank is a transfer-table track and just beyond the transfer-pit are tail-tracks matching the bridge tracks. The unloading bridge, for example, is nm over the transfer table upon its tail-track. The transfer table is then moved to match' the excavator bridge, which is then run upon it, moved over to the other set of tracks and set in place ready for unloading a tank. The unloading bridge in iia turn is put on the transfer table, transferred and set in place on the other set of tracks. SO2 Towers.— In the electro-dcpoaition of copper from a sulphuric add solution any iron in ferric form will be reduced by the current to ferrous form, thus using up electric energy. To overcome this SO2 gas was employed to bring the ferric iron into ferrous form according to the equation
Where care was taken to send to the SO2 towers neutral or slightly acid solutions, this proved easy.
Roastiog for S02> — Referring again to the flow-sheet, Fig. 238, there ate six seven-hearth roasters which carefully roast 75 tons daily of Bisbee pyrite ore. The strong gas, containing 8 to 10 per cent SO3, leaving the roasters passes to a Cottrell precipitator or treater, where it is cleaned from duet before it enters the spray or cooling chamber. Upon the top and sides of the chamber are nozzles by which about 100 gal. per minute of " neutral advance " solution is sprayed to cool the gas before it enters the reducing towers. The ferric iron in this solution is at once reduced to ferrous form.
There are six towers built of sheet-lead and arranged in pairs. They are 20 to 28 ft. diameter by 40 ft. high and rest on a concrete base, the lower edges dipping into a lead-lined sump, 6 ft. deep, so that the edge below the surface of the liquid in the pan forms a seal against the escape of the gas. The tower space is fiUed with boards cross-piled on edge and 1 in. apart. The solution comes to the top of the tower by launder and enters it through gas seals, that is a siphon-shaped pipe, that lets the solution through without letting the gas escape. The solution goii^ tlirou^ these seals trickles down through the boards, wetting them, and is subject to intimate contact with the strong gas, cooling it to atmospheric temperature, and at the same time reducing all the iron present to ferrous form. Between the second and third towers is a 60-in. suction-fan made of lead. This draws the gas from the roasters through the Cottrell treater, spray-
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The Ajo Process 437
cbtunber, and third set of towers, and forces it through the second and firet sets to the atmosphere.
The solution or neutral advance, say 900 gal. per minute, travels countercurrent to the flow of gas, that is, the most reduced solution comes in contact with the strongest gas. The solution from the newest tank of ore is pumped to the top of the third pair of towers (Ss, S3) by a 9-in. centrifugal pump. From the bottom of these it is lifted to the top of Sa, 82, and from their pumps to the top of S, S. i'rom the bottom of these it ifi pumped by Ps to the settling tank, the fourth tank of the nearest row. The purpose of this tank is twofold ; to settle out the slime and to cause additional reduction as the solution stands. The average of ferric iron in the solution entering the toweis is 0.80 per cent and of that leaving the settling tank to go to the electrolytic tank house only 0.10 per cent FejOa.
Electrolytic Depositkni.^ — ^The electrolytic tank house is arranged much as in electrolytic copper refining, which see. The tanks are arranged in banks with silk between. There are 12 banks of 10 tanks each and 4 banks of 8 tanks each. Each tank has 84 anodes and 77 cathodes per tank. The anodes are of hard lead, containing 3.5 per cent antimony; the cathodes are copper starting-sheets, originally 15 to IS lb. in weight, while the finished cathodes weigh 130 to 140 lb. each. Of the total of 152 tanks, 25 are reserved, being employed making starting-sheets, the renuuning 127 tanks are depositing copper on the sheets for the production of cathodes. The starting blanks upon which the starting sheets are made are of antimonial or hard lead like the anodes.
The electric current is passed through each tank in parallel and through them all in series. It is suppUed by two identical units each of 15,000 amperes and each having seventy-six tanks in series.
The Kectrolyte. — The solution entering the tank-house contains Cu 3.0 per cent, Fe in ferrous form 2-3 per cent, Fe in ferric form 0.085 per cent. When it leaves the tank house to go to T. H. R. (tank house return) sump its copper has fallen to 2.5 per cent, that is, 0.50 per cent has been deposited, the ferrous iron has decreased to 1.66 per cent and the ferric iron correspondingly increased to 0.75 per cent. Sulphuric acid has also been set free, there being 2.10 per cent as ag^st 1.70 per cent entering the tank house. The average weight of copper deposited per kw. hour may be given at 0.8 lb.
As will be seen in the flow-sheet the tank-house retum-eolution is made up by addition of solution from tanks E and A and, if necessary, the acid in it is strengthened by the addition of strong acid made at another plant. This is in order that the strongest acid solution shall go to the newest tank. The new acid, that brought from the outside, will amount to 60 per cent of the total acid consumed. Of the remaining 40 per cent some 32 per cent is r^enerated in the S02 towers and 8 per cent at the electrolytic tanks.
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438 The Hydrometallurgy Of Copper
Only about half of the total acid used in an ^ghlxlay leaching is utilized in dissolving copper, the remainder is xised up in dissolving ijnpuiities. Therefore these impurities gnulually accumulate, mulHtig the solution more and more sluggisb. To keep the solution active a portion must be discarded and replaced by fresh water introduced at the laet wash. This daily discard must be such that its impurities shall be equivalent to those taken up each day. In this particular plant it amounts to 90 gal. per minute out of the 1324 of neutral advance.
The copper in the discarded solution is recovered by paamng thioU|^ scrap^ion boxes or launders, just as described for the Rio Tinto process, which see. It takes 2 lb. of iron scrap to precipitate one of copper. The resultant precipitate, which contains perhaps 75 per cent copper, is shipped away for smelting. About 10 to 15 tons of copper is recovered in this way daily. The cathodes from the electrolytic tank house are quite pure, containing as high as 99.85 per cent copper.
AMHONU LEACHraO Umitations of &e Process. — This is suited to the treatment ctf a mixture of sulphide and oxidized ores, that may contain so much calcite aad dolomite as to forbid acid leaching. The sulphides are removed by concentration and the tailing, containing the copper carbonates and oxides, are leached out with a carbonate of ammonia solution. Where, as in the Lake Superior region, the copper occurs native, then this is likewise removed by concentration except the finest partictes of native copper which, with any copper oxide, may be dissolved by the ammonium carbonate. In the case r4 malachite we have the reaction
(12) CuCO2Cu(0Ha)+2(NH4)2CO3-=CuC03, 2NH3-|-HaO-|-2C02,
the CO2 gas escaping freely. The reaction with azurite is similar. The fine copper particles with free access of air are oxidised to copper oxide and
Where the air does not have free access, the cupric salt thus formed comes in contact with the oxidized copper particles and is reduced to the cuprous state.
However, this latter compound, coming where there is access of air, again rapidly oxidizes to cupric form, and can then proceed to dissolve an additional quantity of native copper.
Upon subjecting the solution to boiling by steam, the ammonia and carbon dioxide of the cupric carbonate are distilled off thus:
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Ammonu Leaching
The copper oxide falls out of the solution as a heavy powder, and the ammonium cartxniate is abeorbed by water and this recovered fw further
Ahhonu Leachino At Kbnnicott
We present in Fig. 240 the plans and elevations of an ammoniarleaching plant for the treatment of tailings c<staining 1.46 per cent copper or of 1.14 per cent copper in soluble form as copper carbonates. After ammonia treatment the residue, going to waste, still retained 0.26 per cent of copper indicating a recovery of 77 per cent of the soluble portion.
Referring to Fig. 240, section Z, Z, the tailings are brought by a launder to the 275-toD storage bin. Thence, when needed to fill the leaching tanks (see the elevation), they are raised by a vertical belt elevator, and after a Vezin sampler has taken out a portion, are transferred by a short, inclined conveyor to the long distributing conveyor, seen in the plan running over the leaching tanks. When leached, the exhausted tailings are removed by a taihngs-discharge conveyor beneath. The tanks are 30 ft. diameter and will hold 500 tons. They are filled from the distributing ctmveyor by a revolving plate provided to distribute the feed evenly over the whole area and to give a uniform mixture of coarse and fine. A dome-shaped cover is then put on.
The first leaching solution, consisting of rich copper-ammonia solution from a previous leach, together with concentrated ammonia from the still is run into the tank at the bottom,, and rises through the charge, displacing the moisture of the taiUngs before it. The first part of this moisture, practically barren, is sent to waste. Rich solution is now brought on top of the charge and leaches downward. As fast as the downward flow continues it is returned fron below by means of a centrifugal pump to the top of the charge, this being kept up as long as the solution will dissolve the copper, or for about thirty hours. The leaching solution, carrying 4} per cent Cu and 7.5 per cent NHa, is now drawn off, part going to the evaporators B, part to the rich solution storage tank A. This is followed by a weak copper ammonia solution as a preliminary wash, which follows the rich solution down through the chai^. When this begins to appear it is turned into the wash-solution storage tank A. Following comes the steam wash. Steam, at a pressure of 5 lb. per square inch, is admitted above, and graduaUy works to the bottom, heating the charge and condensing to a film of water which displaces most of the copper-ammonia solution. The remainder of the ammonia, volatilized by the steam, is carried out with it to be condensed in one of the condensers D. When the ammonia content of the vapors issuing from the tank has fallen to 0.5 NHs, the wash is considered finished, the steam is shut off and the tank is emptied of the exhausted tailings. The steam washing takes about twenty
The Hydrometallurgy Of Qopper
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Ammonia Leaching 441
hours, aod about 100 lb. of steam ia \ised per ton of charge. By this method of steam washing one can leach rich ores with little loss of ammonia in the tulings and without using a large volume of wash-water, which would too greatly dilute the distilling solution.
There are three unite of evaporators. Section Y, Y sbowa one of the«e units, the evaporators B B in two stages, each provided below with a filter C. As the copper-ammonia-bearing solution is evaporated, the volatile ammonia vapor paaees over to be condensed in the condenser D. By thus arranging the evaporators in double effect the ammonia, concentrated to at least 18 per cent NH3, is obtained. The copper oxide falls out of the solution as the ammonia leaves it, to the conical bottom, of the evaporator and passes to the filters C, C. The precipitate contains as much as 80 per cent Cu, the filtrate carries but 0.4 per cent Cu and 0.9 per cent NH3. This is re-treated in a secondary operation to precipitate the balance of the copper and to produce a waste solution of but 0.025 per cent NH3.
The ammonia is originally purehased as aqua anmionia or ammonia hydrate, but soon picks up CO3, then becoming ammonium carbonate. Due to the large amount of COa evolved in the leaching, vents are provided in the t&nk covers.
It will be noticed that if the tailings contain gold or mlver in such quantity as will justify it, these metals may be readily cyanided. The tailings^ add free, are in excellent condition for such treatment.
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CHAPTER XXXin
Refining Of Buster-Copper
Copper Repining
Blister-copper, or black copper, whether produced in the blastfuiuace, the converter, or the reverberatory furnace, or by melting tbe concentrate from tbe native copper ore of the Lake Superior region, still contains impurity, principally araenic with sulphur and iron, and all impurity must be removed by refining. If the copper contains gold and ^Iver in quantity to warrant (20 to 40 oz. silver per ton), it is melted without attempting to refine, and cast into anodes that then ate subjected to electrolytic refining. If there is but little precious metal in the copper, it may be direcUy refined in the copper refining-fumace. Fig. 241 is a sectional elevation and plan of a 40,000 to 60,000-lb. copper refining furnace. It is 14 by 19 ft. hearth dimensionfi, and has a firebox 5^ by 6i ft., or 30 ft. area, and carries a fire-bed 4) ft. thick. The hearth, 2i ft. deep, has a brick or a sand bottom. If of sand, the bottom is c&icfully Bmelted in. Beneath, the hearth is vaulted for ventilaticai. The bridge, 5 ft wide, is strengthened by a double conker-plate, and on either rade and in tiie tooS over the fire-bridge, are ports that are opened when an oxidizing flame is demred. In the elevation, at the front end, is to be seen the outlet-flue that leads to tbe stack or chimney. The chimney is close to the furnace, but is not shown in the plan. The charge oi ingots of blister-copper is put in at the side door. The door is then ti^itly closed, and vigorous firing follows. The chaige melts after several hours. The front door b then opened, and whatever slag has formed during the melting is skimmed.
Next follows the rabbling, tbe object of which is to oxidise a portion of the copper and the impurities with it. The operation years ago consisted in striking the surface of the bath with a rabble in such a way as to splash the metal and agitate it, thus exposing it to the action of the air. The present way is to insert a J-in. pipe just beneath the surface ci the metal and force compressed air tiirough it to ^tate, and at the same time to oxidise it. The air-ports of the furnace also are opened and the flame is made an oxidizing one. Tbe action proceeds to the stage of " set copper," CuaO having been by this time formed, and in part dissolved in the copper.
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Copper-Refining Furnace 443
Iron, sulphxir, and arsenic partly volatilize, and partly oxidize and enter the slag that b formed at the same time ; this is skimmed off.
The copper oxide must be removed by poling. This is a reducing action in which the sir-porte are closed to give a reducing flame, and spruce or
poplar poles are inserted at the front door into the metal. The outer end of the pole is raised to force the butt-end beneath the surface of the metal. At the same time a wheelbarrowload of charcoal is thrown in to cover the surface, to exclude air. and to reduce cuprous oxide. As the hydrocarbon
of the wood is evolved and the moisture evaporates, that is, as the wood bums, reduction takes place. The operation requires an hour or two. Additional poles are inserted to replace those consumed. Samples of a few ounces of the copper are removed in a small ladle from time to time and examined to note the progress of reduction. The "tou^ pitch" (the point at which the cuprous oxide is completely reduced to metallic copper) is the end in view.
The charge is now ready for dipjnng or ladling. Hand-ladles, holding 25-lb., or large " bull-ladles " holding 200 lb. and carried by an overhead trolley or crawl, are used. The dipping or molding consumes three hours. . The copper b kept hot by occasional firing, and by keeping the surface of the metal covered with charcoal. The charcoal seires also to keep tbe copper in pitch, or in the condition of toi^;h copper. The molds into which the copper is poiu^ from the ladles are of the shape required by the trade. There are required ingots or bars suited to remelting for making brass; wire bars of a form convenient for rolling into wire; and rectangular c^es, cftoi IS in. square and 4 in. thick, but also of dimensions giving 2000 to 4000 lb. weight. The size of the cakeg is suited to the size of the sheets of copper into which they are to be rolled.
llELTnTG AND REFinma " LAKE " COPPER
The product from which copper is made in the Lake Superior region is a concentrate (called locally " mineral "), which averages 70 per cent copper in native fonn, accompanied with a self-fluxing or fusible gatq^e. In addition there occur pieces of copper of different siaes, from that of tbe fist to several tons in weight, called mass-copper. The small pieces are handled easily, and are shipped to the smelting works in barrels. It is called barrel-work. The larger pieces called " mass copper " or simply " mass," are 70 per cent copper. For the large pieces that cannot be charged at the side doors, a hatch-opening with a clamped brick cover is provided. Large pieces are raised by a crane and charged through the hatch, also concentrate or mineral to make a charge of 36,000 to 40,000 lb. The charging takes place inunediately after the dipping and the repairing or fettling the furnace. The furnace is now closed, and firing proceeds for several hcurs. As the charge melts and slag forms, it is dcimmed until the metal is completely melted and the surface is clear.. The operation of refining then continues as has been described above.. The slag contains 15 to 25 per cent copper, partly as entrained prills and flakes, and partly as cuprous oxide.
This slag is smelted in a blast-furnace with added limestone to make the resulting sl^ fusible, using anthracite coal and a portion of coke for fuel.
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Copper-Casting Machine
In recent practice in the LAke~ Superior country the operation of melting is performed in one furnace and refining in another. A furnace, 18 by 40-ft. hearth area, melts 100 tona of mineral of 67 per cent copper in twenty-four hours, iiaing 30 tons of coal. The charge to the Bec<md fur^
Fio. 242.— Endless Mold CMting Machine.
Pio 243 — Endless Mold OaatiDg Mochue.
nace is supplied in two portions, the second following as soon as the first is melted.
The copper is poled, and when thus refined, is cast into ingots on an endless-mold casting-machine, see Figs. 242 and 243, or on a Walker casting-machine, see Fig. 245. The copper is tapped to casting-machine over a spout which is hinged, the lip being raised or lowered to regulate the flow.
446 REFINING OF BLISTER COPraR
In Rg. 147, the endless chain receives the molten copper upon the chjun of molds at a, and this slowly travels aloi^ until the now solid ingot drops upon a grating and then on an endless chain conveyor in the water box at c. Here it is thoroughly cooled, lifted to d, and drope upon the floor at the foot (tf the slide e to be removed for shipment.
THE UAKHfG OF ANODES AlfD OP COHHERCIAL CATHODE COPPER
Melting Furnace. — The melting down of blister-copper in ingot form is performed in a coal-fired teverberatory furnace, 40 ft. long of a capacity of as much as 400,000 lb.
Charging Machine. — This, Fig. 244, consists of a crane carrying a transverse carriage whereby an arm or ram is introduced into the furnace, the paddle B at the end carrying it« load of several ingots piled on one another. Upon the arm rests a racked bar that receives as inde-
Fio. 244. — Clarke and Antisell ChargiDg-crane.
pendent movement by means of the cylinder c. This, at the ri^ moment, shoves the load off the paddle to drop upon the furnace bottom.
The charge is melted and poled to ^ve a smooth anode, then tapped, in regulated flow.
The duty imposed upon these anode casting-furnaces is extremely severe and constant. They hold about 200,000 lb. of metal, and no sooner has the last anode of a chai^ been cast than a fresh charge be^ns. In order to complete their duty of 300,000 lb. copper (Ij charges per twenty-four hours), it is often necessary to begin blowing compressed air into the metal as soon as the bath is sufficiently deep to submei^ the air-pipes, taking care, however, not to let the oxidation ■ proceed too far before the entire chai^ te collected, else there is danger of such a violent evolution of SO2 gas, from the reaction between cuprous oxide and the cuprous sulphide of the fresh converter copper, that there is danger of the melted metal being blown out of the furnace. Two of these furnaces are in constant use, with a third and larger one in reserve.
WALKER CASTING MACfflNE 447
The pigB contain about 9S.3 per cent copper, and this process brings them up to about 99.3 per cent, at which stage they will cast into smooth anodee.
Pio. 245. — Walker CastitiR Machine.
The Walker Casting Machine.— This, as shown in Fig. 245, is of the horizontal wheel type and is capable of casting 25 tons per hour. The metal Bows tbxougb the tapping-slit at the front end of the furnace into a
suspended ladle B, from which it is poured into an anode^nold attached to a platform conveyor operated hydraiilically. When the mt^ is filled, the ladle m dropped to the horizontal [wattioQ, and ibs coDv^or is moved so as to bring the next mold into position. The copper is chiUed by a ^ray and when " set " is dumped automatically hvm the mold onto a conveyor operating through a tank of water. {
In the case of cathodes from the tank house of a refinery, these are \ charged by the charging machine, Pig. '244, to a large reverberatory fur- ! nace, melt«d down and tapped to the Walker casting-machine. The mokb, as shown in Fig. 24S, are making ingots for commercial use.
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CHAPTER XXXIV ELECTROLYTIC COPPER REFINIHO ELBCTROLTTIC COPPEE-REFIKIHG PLANT
Blister-copper from copper reduction worka in the Western States, as well as that imported, is profitably treated by this process. Not only can pure copper be produced from impure material, but the gold and
Fig. 246. — Ground Plan of Eleptrolytic Copper Refinery.
silver in the blister can be separated, parted and refined. The copper from the reduction works, in the form of rough ingots, or even as anodes, b sampled and assayed to determine the content of precious metal and of copper.
Reflneiy. — ^The following is the description of a refinery of ihe capadty
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Electrolytic Copper Refining
of 60,000 tons per Eomum, and covering 22 acres. It combines the best oS the standard methods.
Operation. — Referring to 1%. 246, the blister-copper is received over the railroad spur B, weighed in the control weighing room E, sampled is the sampling room F, and then placed in storage under the crane-way L, from which it is taken as required, to the furnace building J,K. This has at one end two reverberatory anode furnaces for melting the blister, and two where the cathodes produced in the process are remelted into
¥i<i. 247.— Electrolytic Tank.
merchantable form. The anode furnaces are Uned with siUca brick and are each of 100 tons' daily capacity. The furnaces are fired with slack coal, and have a waste-heat boiler, so as to utilize the steam for power purposes about the plant. The raw material is charged into the furnace on one mde from the industrial railway by means of a charging machine (Fig. 244) and the molten copper is tapped from the other ^de into anode molds. The depth of the bath of molten metal is from 21 to 24 in. and during the twenty-four how period the cycle of operations would be: For charging, 1^ horns; for melting down, nine hours; for refining and
Electrolytic Copper Refining
poling, seven hours; for casting, 6J hours. The anodes are stored under the craneway 0, in plentiful supply to feed the tank-house U, where the electrolytic copper is produced.
The tank-house contains 512 electrolytic tanks in two bays, a bay having 8 sections of thirby-two tanks each. Each tank, to hold 28 anodes and 29 cathodes, is 13 ft. long, and of cross-section shown in Fig. 247.
The current passes through each of the 30 plates or anodes to the cathodes placed between, and anodes and cathodes are 2.6 in. apart. Assuming that the anodes are 2 by 3 ft. in size, we have, in each tank, a total area of 360 sq. ft. through which the current is paaeii^ with a density of 20 amperes per square foot. We thus have a total of 7200 amperes, The pressure in passing through this 2,6 in. of electrolyte to the cathode is 0.40 volt. The tanks in series would therefore give a pressure of 102 volts.
i
Fig. 248.— Current-flow Walker Multiple Syateni,
Fig. 248 represents the arrangement of anodes and cathodes in the Walker multiple system. It will be seen that the flow of current through each vat is in parallel but from vat to vat in series.
The buHdtng must be kept at a uniform temperature of about 80° F., causing a tendency in cold weather to " sweating " on roof and walls, on account of the large amount of moisture evaporated from the electrolytic tanks. For this reason the heating should preferably be by the circulation of dry hot air to absorb the evaporated moisture.
The coiner electrolyte consists of about 4 per cent copper in the form of sulphate, together with about 12 per cent of free acid. During the deposition of the copper the electrolyte becomes, in course of time, polluted by the impurities contained in the anode copper, on account of which a certain amount of the electrolyte is periodically drawn off and treated in the regenerating plant, the purified solution being returned to the copper electrolytic tanks. The rate of decoration of copper on the cathode will be governed
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452 ELECTROLYTig COPPER REFINING
by the current denaty, which should not amount to more than 20 amperes per square foot of anode surface for copper carrying up to 100 oz. of gold and silver to the ton. Should the amount of precious metals be apptBciably less, the current density may be increased up to, say, 30 amperes, thereby increasing the speed of deposition and ^ortening the time oi operation. Should the decoration take place too quickly, there is danger of occluding precious metals with the cathode copper. A cycle of opetstions in the tanks at the 20-ampere current denedty would occupy from three to foxir weeks.
Starting Sheets. — One section of the tank house is devoted to the preparation of cathode starting sheets, for which purpose a copper plate coated on each mde with oil or graphite is used as a cathode blank, and a thin layeaof copper deposited thereon by electrolysis. This is stripped off from each side and forms the cathode starting sheet. 'In the routine operation the electrolytic tanks are first chai^d with anodes, and then himg with cathode starting sheets, and the current is started. It is kept on for a prescribed number of days, varying from ten to twelve; then the cathodea which have been formed on the starting sheets are removed, the solution is lowered in the tanks, the anodes are taken out temporarily, and iho slime is removed and sent to the slinie refinery. The anodes are then replaced, fresh cathode starting sheets hung, and Hie current is turned od for another period of ten to twelve days; then the " pullii^ " is npeated, except that upon this occaaon what is left of the anodes is removed and goes back as scrap to be remelted in the anode furnaces. It will be noted, therefore, that for each anode going to the tanks two cathodes of lighter weight are fonned. The theoretical deposition of 1^ copper would be approximately 0.062 lb. of copper per ampere day, and a current efficiency of at least 90 per cent should be acquired. The lecessaiy pressure would be about 0.4 volt per tank, and upon this baas the rate of deposit should amount to about 6 lb. of cathode copper per kilowatt-hour.
The cathodes are weighed on leaving the department, and stored, from which point they are fed to the refined copper furnace for melting into the shape required by the consumer, for example, ingots, wire bars, and cakes.
The Slime Refinery W, Fig. 216.— The precious-metal slime reclaimed from the electrolytic tanks amounts to about 30 tons per month, containing about 440,000 oz. of gold and silver. . The slime is first freed of t^ electrolyte by settling until it contains about 50 per cent moisture. It is then boiled with sulphiuic acid to remove soluble copper, after which it is put through a filter press and roasted. This roasted product is combined with a portion of xmroasted slime to make up a furnace charge. The result will be the melting of about 20 tons per month of slime containing frmn 15 to 25 per cent moisture, the melt taking place in a small reverberatory furnace, the product of which is dor6 bars.
Handling Of Anodes And Cathodes 453
Parting. — Dor^ bars form the anodes for an electrolytic deposition, which takes place in an electrolyte slightly acidified with nitric acid, the ailver being deposited on cathode blanks in crystalline form, and the gold settling as a black mud containing about equal quantities of gold and silver. The resulting ailver ciystals are scraped off the cathode blanks, washed and melted into fine silver bars. The gold mud is purified by ticatment with nitric acid to part the gold from the silver, the solution going back to the silver electrolytic tanks, and the go'd sand melted with borax into fine sold bars.
Fig. 240.— Mechanical HaocUing of Anodes and Cathodes.
Mechanical Handling of Anodes and Cathodes. — In Fig. 249, the view above shows the anodes, and below, the cathodes, the entire anode or cathode contents of a tank being lifted and transferred in one operation. The frame shown is brought down by a traveling crane over the tank, all the anodes or all the cathodes of the tank hooked on to it, the frame and its load lifted and transfeixed to a clear space at the end cd the tank house where the fragments of anodes can be removed to be sent to the anode furnace, while the built-up cathodes are stored, to go later to the refinedHx>pper furnace.
The Regenerating Haut V, Pjg. 246. — The best proportion of acid and
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454 Electeolytic Copper Refining
copper for the electrolyte is 10 per cent H1SO4, and IS per cent CtiSQi (eqiiivalent to 6 per cent Cu). When the copper exceeds this quantity, the resistance incieases; hence copper is removed from the circulating electrolyte or solution if in excess, to brii^ the amount to the required proportion, the quantity (tf iron, arsenic, antimony, and telhirium gradually increases, and a time comes when the electrolyte becomes foul with them and the excess must be removed. Antimony can be kept low by the daily addition of a small amount of salt, which fn^cipitates as an oxychloride.
To purify the electrolyte the following method is used. A portion of the electrolyte is diverted in a constant flow to tanks reserved for the purpose of purification. These have insoluble sheet-lead anodes and copper cathodes. A strong curront is used, bo that not only is cojiqier deposited, but also the impurity. The deposit collects loosely upon the copper plates and falls to the bottom of the tanks. Every two months tiie accumulated mud, containing 40 to 60 per cent copper, is cleaned out and reduced in a reverberatory-^defining furnace to form impure bars of copper. The purified electrolyte is returned to the main system.
Circulation of the Electrolyte. — ^To avoid short-circuiting, and to increase the activity and regularity of depoation, the electrolyte is made to flow or circulate through the tanks, entering the top of each tank near the ead, pasang downward between the plates, and finally ri£dng and flowing away through an overflow pipe at the other end. After the sohition has flowed throi^ two tanks in this way, it enters a laimder that returns it to the collecting or Buni{>-tauk. Thus every pair of tanks has an independent circulation. The sump-tank receives all the electrolyte, to be here heated by means of a steam-coil to 40** C, the effect of the warming being to decrease the electrolytic redstance. It is then pumped up to a distributing or stock-tank, and once moi« enters the circulation.
Testing the Current — Besides the voltmeter and ammeter to be found at the switch-board in the power-house, it is custfHuary to use a voltmeter for constantly testing the drop in potential between the anodes and cathodes. For this a forked rod is used, which touches the two plates and takes a small current through a portable voltmeter. A slight drop of pressure indicates short-circuiting.
The Power-house, Q, Fig. 246. — Here are installed five motor generator sets each to give a constant current of 5000 amperes at 60 to 115 volts. Four of these are arranged to work in groups of two in parallel directly to the two separate circuits in the tank-house. The fifth tint is connected to act as a spare for any of the others.
Catiiode Storage. — Leaving the tank house as heavy cathodee, these are placed under the craneway 0, thence dehvered to the reverberatory refining furnaces J, there to be melted and cast in the forma required by the c
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REFININO AND OFERATINa COSTS
CAPITAL RSQUniEUBirrS
Not only is capital invested in the buildings and the equipment of the plant, but it ia required for:
(1) The stock of anodes in process of treatment.
(2) The stock of anodes awaiting treatment.
(3) The copper constantly contained in electrolyte.
(4) The copper needed for the heavy conductors transmitting the current.
The result of this lai^ demand upon capital is to restrict the operation of plants to places near financial centers, like New York, where cheap money is available, the copper near the market, and labor abundant. These conaideratifHiB may outweigh the advantages (^ having the plant near cheap water-power.
COST OF BEFmBRT AlfD OPERATINO COSTS
Cost of Refinery. — ^To refine 50,000 tons of blister-copper annually contuning gold and silver, the output, consisting of refined copper in the shape of wire bars, cakes, and ingot, of gold as fine gold bars, and silver as fine silver bais, the construction costs will be as follows:
Construction Costs
Furnace buiiding t 67,500
Two anode units 95,000
Two refined-copper units 92,200
Four cranes, two aervicea, two charging 40,000
Auxiliary equipment S4,800
Total tuniace department
Tank houae building $127,000
Sixteen 32-tank sectiona and circulation ^atem and equipment 258,000
Electrolytic circuit conductors, etc 90,000
Four service cranes 25,000
Auxiliary apparatus 51,000
Total tank house department
Fower-bouse building and crane $ 42,000
Five electrolytic motor generator seta and equipment 116,000
Motor-driven pumpa, etc 32,200
Auxiliary apparatus. , 30,200
Total power-house department..
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Electrolytic Copper Refining
Slime refinery buildiiig t 38,000
Settliiit! and boiling tanks, filt«r prcM and niutN 16,S00
Electrolytic partbg cells and equipment 18,800
Gold-refining equipment 2,600
Auxiliary apparatus 11,700
Total slime-refining departmrait S10B,700
Shops building and equipment S 31,000
0£Sce and laboratory and warehouse equipment 44,000
Sampling apparatus 7,S00
Storage cranee, craneways, and locomotive crane 36,700
Receiving and shipping apparatus 3,000
Industrial locomotivee and cars 47,800
Industrial tracks and railroad sidings 24,S00
Miscelianeons auxiliary apparatus 7,000
Bluestone and acid neceesary to make up electrolyte 25,000
Total miflceUaneoua S23S,10D
Total $1,502,800
These figures are the more valuable that they show in what unusual ways the money has to be spent, causing costs to mount rapidly.
Operating Costs. — These are as below estimated, viz. :
Labor.
Total Per Ton.
I1.15S 2,025 1-00 .463 .926
S5.66g
5Chbdulb Op Copper And Copper Ore Fbicbs
In Utah, custom copper smelting works pays $10 an ounce for the gold; for 95 per cent of the ailver, based on the New York quotation, and for copper, after deducting I per cent from the wet assay, a further deduction of 3 to 4 per cent from the New York price per ounce. A charge of %5 per ton is made for treatment and when the insoluble exceeds 40 per cent then 5 cents per unit for all over this. Zinc in excess of 10 per cent is charged for at the rate of 30 cents per unit. The deduction of 3 or 4 cents from the copper price is to cover the cost of freight and refining.
Prices Of Copper And Copper Ores 457
Copper. — The New York price is expresBed in cents per pound, quotationa beiag ipven tw Lake copper cast in the form of cakes for rolling into sheets, into ingots for remelting to make castiogB, brass, and bronze, or for wire-bars for drawing into wire. Casting-copper is not as pure as that which is to be rolled into sheets or drawn into wire. Electrolytic copper is made by remelting cathodes (the product of electrolytic refining) into ingots, cakes, or wire-bars. Cathodes aic held at ) cent leee than electrolytic copper, the diffcRnce paying for the remelting.
In London copper is sold by the long ton in English money, and is of various brands. Standard copper, formerly called g.m.b. (good merchant^ able bars), is the grade upon which the others depend. Besides these brands we have :
" English tough copper," " best selected " or " standard." If sold for immediate delivery it is called " spot copper"; if the customer wiU take it at tile expiration of three months it is called " three-months copper."
It is the business of dealers, and others interested in copper, to keep statistics of the supply of available copper, which is called the " visible supply." When this is small the price naturally riaee, and the mvetae is true when it is large.
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Part Vi Lead
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Chapter Xxxv
Properties Of Lead And Its Ores
n^sical Properties of Lead. — ^This metal, bo soft that its purity may be judged of by scratching with the thumb nail, is commercially divided into corroding, common, and antimonial lead. Of the base meals it is the heaviest, having a specific gravity of 11.37. It soon tarnishes on a freshly cut surface, becoming dull gray. It is malleable, so that it can be rolled into sheets and pressed into pipe, but it has little tenacity. It freezes at 325" C. and boils at 1525° C.
Characteristics Of Lead Ores
Classes of Lead Ores. — The lead ores are those in which lead is the principal constituent. The term b applied also to mineral ^gregates consisting of jnore than 10 per cent lead. The lead ores may be divided into two classes, the sulphide and the oxidized. The terms are used only according to the constituent that is in excess; in many lead ores both sulphides and oxides are found. Ore containing no lead is called dry, and when carrying lead, leady. The tatter term ia the oppoate of dry, but we do not term a leady ore a wet one.
Galena. — Pure galena contaii^ 86.6 per cent lead and 13.4 per cent sulphur. In nature it occurs with gangue or vein-matter. When there is much erf the latter it can readily be concentrated. The following table gives an ides of the lead-content of ore, before and after dressing:
Galena Or£S
Co«.«™™.
Pb.
Pb. Per Cent.
S. E. MiflBouri
Minnie Moore, Wood River, Idaho
St. Joseph, Mo
Kellogg, Idaho
Col. Sellere, Leadville, Colo
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Properties Of Lead And Its Ores
GaLena from the Mississippi Valley contains little silver. Galena conoraitrate from S. E. Missouri contains Pb, 69 per cent; SiOg, 1.4 per c«it; Fe, 5.1 per cent; CaO, 3 per cent; Zn, 0.8 per cent; S, 15.5 per cent. That from the Rocky Mountain r^on is not only argentiferous, but nmy contain gold. The precious metals as well as the lead detennine the vahie. Metallic sulphides, such as pyrite and blende, are often aasociated with galena, and with the gangue may carry so much of the gtJd and silver that concentrating leads to a serious loss of the metals and is (Hnitted. If by hand-picking ore can be brought to contain 30 to 40 per cent lead, it is a desirable ore for the smelter. When of this tenor in lead, and free frcxn other sulphides, it carries bat 5 per cent sulphur and needs no preliminary roasting, and is smelted directly.
Oxidized Lead Ores. — Little lead oxide is found in nature. The ores classed here imder oxidised ores are the result of the alteration of galena. They inchide the carbonate (cerussite) and the sulphate (anglesite) of lead. The minerals are mixed with metallic oxides and vein-matter or gangue in nature, and when sandy or earthy, the ore is called sand or soft carbonate, and when hard and stony, hard carbonate. In many deposits we find ore that originally was galena, profoundly altered to cerussite or aDgleait«. The subjoined table gives the compomtion of some of the so-called carbonates:
Carbonate Orbs
Loealily.
Par Cent.
SiOb Percent.
CaO, PwCent.
O.fl l.S 2.0 6.0 8.3
LaftdviUe Colo
Eureka, Nev
Hom Silver mine, PriBco,Utab
Of the ores of the table, that from Eureka, Nev., contains 4.2 per cent of arsenic, which forma an arsenical speiss when smelted. The Hoin Silver ore, apparently oxidized, has the lead in the form on an^eaite (PbSOi), and matte is formed from it in smelting. In oxidised ores the silver is apt to occur as a chloride; the gold probably is native.
There are many lead minerals, but those not mentioned occur in small quantity and are not considered among the commercial lead ores.
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Smelting On The Ore-Hearth 463
When lead concentrates are to be smelted only for the lead content, as is done in parts of the Mississippi Valley, a simple plant with a reverberatory furnace,* or the American ore-hearth, is sufGdent. In the rocky Mountain region the lead ore is not smelted to recover only the lead. The lead of ihe ore is employed as a collector of the gold and silver of other ores that are smelted at the same time. By use of the ore hearth, a large part of the lead is recovered cheaply and simply, but a part ia loet in the resultant slag. In silver-lead smelting it is essential that the slag be comparatively free from lead and consequently from silver.
Shelting On The Ore-Hearth
The oie-hearth cannot, as regards capacity, or cost per ton of ore treated, be used in silver-lead smelting. It can be quickly started or stopped, and put in operation with little cost for fuel, so it well serves the purpose of extracting the lead from small amounts of low-silver ore, from time to time, by the men who themselves have mined the ore.
The HeardL — Fig. 250 represents a sectional elevation, a front eleva^ tion of the lower part, and a plan of an American ore-hearth. It consists (tf a cast-iron pan or crucible a, 2 by 2} ft. by 1 ft. deep, to contain a bath of lead. The back p and the two sides n, n, above the crucible, are water-cooled castings. The blast from a fan-blower {not shown) enters by the tuyere-pipe h through the back at o. At ^ is a sloping cast-iron plate called the work-stone, and at t, a pot, placed to recover the lead that flows down over the work-stone. The pot is kept hot by a wood fire below. The structure is surmounted by a brick top to receive and carry off the fxunes.
Operation. — By means of the blast, a glowing coal fire is made that fjllfi the crucible of the hearth, a; re^due from the previous run, containing metallic lead, and 15 to 20 lb. galena, not finer than pearsize, is spread over the fire. The chai^ soon becomes red hot, and the lead, set free, finds its way to the crucible at the bottom. More ore is then added, and the material in the hearth is pried up gently with a bar to keep the mass open and hot throughout. Lumps form, and are drawn out on the work-etone, g, and gray slag that forms at the same time ie separated and the rich residue returned to the hearth. Ore and fuel are again added,
* The b«»tmeiit of lead ore in reverbeistory funucm has not made headway in the United States. There are two reasons for this: In the silver-lead distriots, the ore has not been of sufiBcient grade in lead to warrant the treatment, and lead ore haa been in great demand as a collector to mix with other ores. Secondly, in the Miastssippi Valley where silver-free lead concentrate is made, the question of skilled labor for reveiberatory furnace woric has bad an influence.
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Properties Of Lead And Its Ores
15 to 20 lb. at a time, and operations continue until lead fills the crucible, while on tbe top floats the fuel, unreduced ore, and balf-fuaed material. One man with a bar at intervals loosens and stirs the charge, raising it slowly, while another with a shovel draws upon the work-stone the half-fused mass floating on the lead. Here he separates and rejects the gray al^, and returns the rich residue to the charge. A fresh charge is then added, and the work progresses in the manner described. The lead over-
Fio. 250. — Americad Ore-hearth.
flows the crucible and runs down a groove made in the work-etone into the kettle, i. When the kettle fills, the lead is skimmed and ladled into molds. With air in excess the red-hot galena is in part oxidized to PbS04, in part to PbO. Both these react upon the galena thus:
PbS+2PbO -3Pb+SOs. PbS+PbS04-2SO!.
Also the Rowing fuel, reacting upon any remaining oxide, completes the reduction.
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The Newnam Hearth
To operate the ore-hearth, a blower and power to run it are needed. Much lead is volatilized, and so the treatment ifl not suited to argentiferous galena. The gray slag that is produced stiU contains 35 to 40 per cent lead, and ia sold to ^nelting-works. The direct recovery of the lead is 75 to 85 per cent, the higher figure having been obtained in recent practice.
The Nevnam Hearth. — A recent development of the ore-hearth has been the introduction of mechanical rabbling, thus doing away with the
moet laborious and hottest work of the hearth, that is, the rabbling of the chai^. The increase in its size and capacity is another advantage.
Fig. 251 shows such a rabbling machine, which is electrically traversed upon an overhead track, free from possible obBtniction, The machine carries a rabble or poker, which plows its way through the crMst of the molten bath from the left to the right end of the furnace, a distance
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466 Properties Of Lead And Its Ores
of 8 feet. Following it is a furnace helper, who, with a lon^handled shovel, removes the gray slag and pushefl back unfused lumps, while the fumaceman adds a thin layer of ore and a little coal, as in the Ametican ore-hearth as just described. The rabble, seen in 251, doping downward into the bath, is now lifted, and the machine returned to the left end of the furnace to make another stroke.
Fig. 252 is a cross-section of the plant with the many details plainly marked, the goose-neck rising above the building carries the dust and fume to a balloon flue, where the material is withdrawn into a car.
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Chapter Xxxvi
Silver-Lead Smelting
8Ilvsr-Lsad Blast-Furkacb Smeltiko
This is a blast-furnace method of treatment, applicable to a great variety of ores containing lead, mlver, gold, and even copper. By it, opes containing the precious metals with no lead, are treated with lead-bearing ores, thus using the lead of one ore as the collector of the gold andasilver of another. This is the most effective method of treating such ores. The precious metals are extracted from the ore by a Uast-funutce treaUnent, uaing lead-bearing ores, carbonaceous fuel, and fiux.
Oxidized ore can be directly smelted in the blast-furnace, but sulphide ore b first roasted. Methods of roasUng ore described in the chapter on Roasting.
The ore to be smelted is chained into the blast-furnace as in iron or copper smelting, with a calculated quantity of flux, which, for lead ore, is iron ore and limestone. The precaution is taken to use lead-bearing ore enough to make the lead content of the charge at least 10 per cent. It has been found that if a smaller proportion of lead than this is used, the precious metals are not so well collected in the base-bullion, or work-lead, produced in the smelting. To a charge, as above constituted, is added 10 per cent or more of coke, not only to melt the charge, but to reduce the lead oxide to metal and the ferric oxide to the ferrous form.
Becbiving, Sampling And Bedding Op Lead Ores
Where ore is treated in a small way for the recovery of the lead, as in Missouri, no particular provision is made for storage. In various custom smelting works in the Rocky Mountain region, where lead ores are treated with others by methods of silver-lead smelting, and where ores are boi^t outright for treatment, the h»tiHtitig becomes comphcated. A plant treating ore from its own mine is caUed a mine's works, and here less attention is pven the sampling and storing of the ore. In a custom works, therefore, ores of many kinds are received, some containing lead, some having little lead but carrying silver and gold. 467
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Bffl>DmG ORES AT A CUSTOM WORKS
The ore is received in lote of a few tons up to those of eeversl cat-loads. Each lot is separately weighed, sampled as described in the chapter on Sampling, then assayed, and purchased. If different kinds of oi« were smelted separately the process would involve endless change and labor, and so it has become the custom to " bed " the ore in large bins or stalls, each holdii^ several hundred tons. When so bedded the mixture, called a " mix " is treated as a mng\e ore The different kinds of
Flo. 254.— Ore-bed.
ore are unloaded separately into the bin, and each kind is spread oiit in an even layer before tlic siiccijeding one is added, as indicated in Fig. 254. When the ore is to be used, shoveling is done at the floor and all parts above fall down and mix, since a steep face of ore is constantly maintained. Thus a uniform mixture of the different ores is obtained for smelting. The contente of the bed are treated in the books of the company as a single ore.
Side Ores. — Besides am bedded in this way, large lots may be kept separate. Even smaller lota, of which a moderate amount is to be used per chai^, may be so kept. Such are called " side ores."
Charging from Feed-bins. — Ore and fluxes are often stored in hopper-bott(HiL bins or pockets, to be drawn oEF in weighed quantity into the flirty
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Bedding Ores 4&9
Dace charge-car. Fig. 253 showa such an installation. The chai^ is dropped into the furnace by opening the double-hii^d bottom of the car. The coke and fluxes are stored separately in large piles, so that in case of failure of r^lroad dehvery due to washouts, etc., the furnace shall not have to close
Charging Fuel. — Coke for the cbai^ is forked into the coke-buggies, using a fork with IJ-in. spaces, thus leaving fines which are generally thrown away. Such loss may amoiint to 5 per cent.
A bed, formed of 10 to 15 per cent SiOa, 20 to 28 per cent Fe, and 20 to 28 per cent Pb, roaste well. Mixtures containing less lead and more pyrite than this roast readily, but a mixtxire pulverulent, when
roasted, tends to make more Fio. 253.— Charging Feed-oar.
flue-dust in the blast-furnace,
while with the proportion of lead above specified it sinters and makes a dedrable lumpy product.
Sintering Ores. — Besides sintering by blast-roasting, the Dwight- Lloyd machine, described on page 112, has come extensively into use, not only for silver-lead ores, but for fine iron ores as well.
GEITERAL ARRAHOEUEnT OF A SMELTING WORKS
A On&4unuice Smelting Plant.— In Fig. 256 we illustrate a completely assembled works. Since the materials pass downward from level to level this is called a side-hill or terraced plant. To suit present practice the reverberatory roasting furnaces would be replaced by a multiple hearth-roaster and a Dwight-Lloyd sintering machine occupying but one-third the space.
The 30-in. Dwight-Lloyd machine has treated 60 tons of sulphide ore, or 80 tons where non-sulphide ores are mixed in. The 42-in. machines will yield 80 to 120 tons under similar circumstances, and at higher speeds 200 tons and over.
Fig. 255 is a cross-section of a Dwight-Lloyd sinter plant, contfuning 42-in. machines. An inclined conveying belt delivers charge to a
honzontal one on the top floor which has a tripper to deliver it to the feed hoppers. At the right it discharges to a raikoad car receiving the sintered ore. From the building a 9-ft. balloon flue leads away the gases to a stack. An exhaust fan, one to each furnace, draws these away to the stack.
C(»ipo8itioa <tf ttie Charge for Skktering. — The sinter chai^ may be quite variable. It is customary to rou^-roast or« high in sulphur until it is reduced to about 12 to 15 per cent in sulphur. A satisfactory charge would contain 13 per cent Fb, 23 per cent Si02, 25 per cent Fe plus Mn, 12 per cent sulphur. This rough-roasted product is blast-roasted down to 3 per cent sulpbiu.
Ore in cars, which has to be sampled or crushed for roasting, enters by the track a' on the extreme left, or otherwise by the track j, Fig. 256. At liie latter track, while unloading, every tenth shovelful can be retained in the
Fio. 255. — Cross-section of Sinter Building.
arc while the nine-tenths is bedded upon the mixing floor. The car can then be sent to the sampling mill by track a'. Crushed ore for the roasters is received in a car h, then trammed to the pile d. Here it is withdrawn, as needed, and taken to the hoppers e', c', of the two reverberatory roasters in the roaster-building H. A flue X, common to them both, leads to the stack /. The roasted ore is taken by wheelbarrows to the cooling floor R, then to any bin V of the mixing floor. Non-roasting, sampled ore is trammed by the tracks «', t, for bedding at any bin V. Fuel and flturos are also unloaded to the mixing floor. Charges, as fast as assembled and weighed, are raised by the " lift elevator " to the blastr-fumace feed-floor. The blast-furnace is supplied with air by a rotary blower w, dhven by a steam engine in the " enjpne-blower room."
The gases from the blast-furnace pass away to the flue p lyy the gooseneck down-take o, thence to the dtist^chamber I and to the stack g. The
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Silver-Lead Smexting Works 471
base-bxillion is loaded into the car n, at the front of the furnace building. The slag, flowing through the fore-hearth m, is taken by alag-pots to the edge of the dump and there emptied.
THE SILVER-I£AB BLAST-FDSITACB This differs from the copper blast-furnace principally in the crucible which in the copper bla^fumace is above the floor level on a carriage while in the silver-lead furnace the crucible reste upon the floor, having a deep cavity, the crucible proper.
Fig. 257 is a perspective view and Fig. 258 ahowB elevations of a furnace, giving many details of construction. The brick cnicitJe, heavily bound with plates and steel mils, has, as shown in the transverse section, a channel called a " lead well," widening upward from the bottom of the cnicible and having a spout where the molten lead (which fills crucible and lead-well alike) may be removed. There are six water-jackets on
FiQ. 257. — Penpective View of Silver-lead Blast Furnace.
each Bide and one at each end (see Fig. 25S), fonning the furnace lK>sh. Another tier of jackets forms the lower part of the shaft, while its upper part of brick reaches to the chai^ door. On each long ade are two counter-weighted doors. When a furnace is to be fed the doors are opened, one at a time, to do so. Kach lower eide jacket is pierced with two tuyere openings, making twenty-four tuyeres in all, there being none in the
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Silver-Lead Blast-Furnace
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end jackets. The jackets are held in place by a frame of I-beam that bear against them. There is a main water-supply pipe whence proceeds a feed-pipe to each jacket. The spill from the jackets is taken to a waste trotigh, set above the bustle-pipe. The bustle-pipe surrounds the furnace on three sides.
The tuyeres themselves are well shown in Fig. 259. Each one has its own Rate-valve so that it can be shut off for any desired purpose.
Air is suppUed by the bustle-pipe e, Fig. 261. Iron-pipe connections are now preferred, as in Fig. 259.
In the case of the cloeed top the gases pass away to the dust-flue p of Fig. 256 by way of the down-take o, which slopes at an angle of 45°;
Back To Remove Jacket
Flu. 2 9. — Blaet-fumare Tuyere.
SO that the flue-dust will not lodge in the pipe. At Fig. 261 the down-take is flatter, but it can be cleared by a" rabble worked from the feed-floor.
The arrangement of the wat«r-cooled tap-jacket, as shown in Fig. 262, is an end jacket, having a bosh. Beneath this is set the casHron water-cooled tap-jacket t, 12 in. square. This is wedged beneath and on either side with brick and fireclay, and the breast is bricked up. The conical aperture or tap-hole is plufo^ with clay. When sufBcient slag has accumulated within the furnace, say every fifteen minutes, a hole is pierced through the clay of the tap-hole and the slag flows out over the spout s, being received in a slag-pot or into a fore-hearth.
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Open-Top Blast-Furnace 476
Open- And Closed-Top Blast-Fuknacbs
In the blast-fumace building, Fig. 256, and in Fig. 257, are views oi a closed-top furnace, while in Fig. 261 we have sections of an open-top one. In the former the smoke is caught in the cloeed-top and passes away overhead and the furnace can be fed to the level of the feed floor; in the second the gases pass away by the down-take x which limits the charge level to its bottom. On the other hand the whole furnace is accessible
Flo. 261— Lead Blast Furnace.
from above both for dumping charges and for cleaning out the furnace after shutting down.
The Open-top Furnace. — The furnace, Fig. 256, is fed by hand, shoveling the charge through the feed-doors into the furnace. At lai^ smelting plants this method has been superseded by " automatic chai^g " (see Fig, 261). To prepare for this the closed top above the feed-floor is omitted. Then for charging a large car having a drop-bottom is brought to the furnace and the contents of the car dropped into it. Such a furnace is shown at Fig. 261. On the left we have a half section, half elevation, transverse to the furnace; at the right there is shown on one half, a loncptudinal section, and beside it a fore-hearth. On the right half
is the longitudinal elevation and the brick down-take x, also a croes flection of the down-take.
The furnace ia 46 by 160 in. at the tuyere level, and widens to 95 by 160 in. at the top. There is a i^gle steel jacket at each end, and two boshed jackets at each side. It will be noticed that the shaft widens out as it ascends. There are twenty tuyeres, ten on each side, the opening in the jackets being 3) in. diameter. E^h tuyere has its own metal blowpipe uid shutr-off valve. The down-take x is 3 ft. 3 in. by 7 ft. 11 in. inside dimensjons, and leads the smoke to a flue common to several furnaces. When a furnace is to be charged a light sheet-iron cover at the floor level is readily removed by hand. At 3 ft. below the charge-floor are set transversely angle-iron " spreaders," so that the materials, fallii^ from the car, shall be spread or distributed evenly. In this particular installation the charge-«ar, of about the same width and length as the fiimacc, approaches it at the side; in other cases at the end of the furnace.
Operating The Blast-Fushace
Blowing In.— 'The crucible and lead-well are dried out and warmed by a wood fire for twenty-four hours. They are then cleaned out and the' crutible is fine<l, using the well-cleaned lead ingots from the dressing kettles. These are piled in cross order, filling the crucible. On them is laid a wood fire as high as the top of the breast, and after closing this the fire is lit at the tuyeres. The wood ablaze, more dry wood is added, then coke to the depth of 2 ft. with a little iron ore and limestone, just enou^ to flux the coke-ash. Then begins the charging, at first using double the ' usual quantity of coke, thcn.with the furnace full, dropping to the usual proportion.
The blast is increased gradually during one to two hours, until the furnace is in full operation. As the slag accumulates it is tapped, while the lead, accumulating in the crucible and lead-well, is either dipped from the lead-well with ladles, or tapped through a lead tap-hole at the le\'el of the top of the crucible. The amount ot lead removed at one time is Umited to 1000 to 1500 lb. to keep the crucible always full. The operation of filling and starting takes seven houra.
Regular Work on tlie Cbai^-floor. — In a small furnace this consists in wheeling ore and fuel from the bins to the charge-scales, wei^iing the required amounts into charge-barro^vs or charge-cars, then dumping the charge thus prepared upon the feed-plates in front of the furnace doors. Every material except the foul slag is weighed, and even the slag is added by shoveling in regular amounts. A charge ia dumped on one chat^e-plate and the coke for it on the other. The coke i:? fed in an even layer, and the plate thus cleared receives an ore-charge. The ore-chai^ on the other
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FRONT OF BLAST-FxmNACE 477
charge-plate is then added, taking care to place the larger ore at the middle of the furnace and the fin^ at the walla and corners. Such a distribution should be made as to cause the smoke and gas to rise evenly throuKhout. The blast tends to ascend at the walls more than in the middle, but by the distribution it is compelled to rise evenly. The plate, now cleared of the ore-chai^, receives the coke that is next to go in. The content of the furnace remains at the level of the charjip-floor, new charges being supplied as the surface sinks.
R^ular W<n'k at ttie Ground or Slag-flooi. — This consists in regulating
FiQ. 262.— Front of Silver-lead Blast J'urnace.
the water-supply at the jackets, seeing that the tuyeres are clean and open, tapping and stopping the slag, and when the slag and matte are separated in a fore-hearth, tapping the matte, placing the slag-pots (see Fig. 262), and removing them to the dump when full.
Front End of Furnace. — We show at Fig. 262 the front end of a large blast-furnace. Immediately at the front is to be seen the fore-hearth, d, 6 ft, by 9 ft. in size, mounted on wheels, and here the separation of the matte from slag is effected. The top of the slag in the fore-hearth soon becomes crusted over with a slag crust, but beneath this is the molten
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slag and matte. The heavier matte settles out while the lighter slag escape:; by a spout at the front of the fore-hearth into the bowl of a large slag-car, g, mounted on wheels. There is a tap-hole on the side opposite that shown, where the fore-hearth is tapped in order to remove the matte which then accumulates. This matte is generally received into Sat or dish-shaped bowls and, when solidified, it is removed by a traveling crane. It is broken by hammers and is then ready for crushing to 3-mesh siae for roasting.
When a slag-bowl is filled it is removed by an electric motor to the dump. A shell or " skull " of solidified slag has formed, lining the interior of the bowl and the top. A hole ia broken in the top and the molten contents pouted out as shown in Fig. 263. This shell is letumed tothe feed-floor to add to the chai^, since it contains enoi^ lead and silver to make this worth while. Of late some smelting works reject it all, saying that it contains so much zinc aa an impurity that it does not pay to return it.
Fic. 263. — SidfMlumping SUg-pot.
Slag is also conveniently removed by means of lai^ slag pots mounted on trucks and drawn by an industrial locomotive, see Fig. 155, where the firct two trucks are carrying shallow matte-bowls, the farther two slag bowls,
Drossing Base-bullion. — When handling base-bullion from the small blast-fumace of 50 to 100 tons capacity of a generation ago, the ordinary custom was to tap it when full from the lead well, into a cooler, a pot of one ton capacity. The dross of the metal would form a crust on top, and by aid of a perforated skimmer was put into molds, the clean lead being then added on top to yield a smooth-looking bar. The refinery got the copper-contaJning dross, and so did not complain.
The present method, using large furnaces, is to dross the lead as follows: The lead is tapped into pots on wheels, and this is accumulated in the dros»- ing-kettle until it is filled with the molten lead. It is now skimmed, using a Howard press, the skimming, after pressing, being dumped upon a chute.
Chemical Reaction Of The Blast-Furnaces 479
where it ia broken up and retTimed to the blast-furnace. The hot molten lead is allowed to cool to the casting temperature and Biphoned into molds, as described elsewhere. The resultant lead, well freed from copper, still retains all its antimony, say 1.5 per cent, and this is sent to the lefineiy. The product is quite eves, and its assay is accurate.
The surface of the chai^ should look dead, showing no visible heat or flame (over-fire). With much over-fire, there is a loss of lead due to volatilization. The moisture in the charge soon dries at the temperature of the ri^g gases (200° C). The heat thus absorbed is small and by calculation it is found to be but one-thirtieth of the total supplied by the fuel when 5 per cent moisture is in the chai^. As the materials of the charge descend in the furnace, carbon dioxide begins to be driven frtnn the limestone, and the iron reduces from the ferric to the ferrous form. Half way down at a temperature of 800° C. the reactions are complete. The lead in oxidized form is reduced by the CO of the gas and by the red-hot coke. Galena reacts with the iron oxide and carbon as follows:
Sometimes scrap iron is added to the charge, and acts with ^ena or other sulphides as in the nail assay for silver, as follows:
As the lead, thus reduced, drops through the charge, it collects the gold and silver as well as a part of the arsenic and antimony, and enters the crucible as base-bulhon. When antimony is present it is reduced like lead and alloys with the latter. Anglesite is reduced by contact with the fuel and iron oxide according to the following reaction :
Where much ai^lesite is in the chai^, more than the usual quantity of fuel is demanded. Since the affinity of sulphur for copper is greater than for iron, copper sulphide remains tmreduced, and copper as an oxide takes sulphur from the charge and with the iron fomas matte. Lead to the extent of 10 to 20 per cent, either as sulphide or in metallic form, is also taken up by the matte. To some extent sine sulphide also enters the matte.
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The ferrous oxide, not needed to satisfy the matte, and the CaO and MgO in the charge must be present in sufficient quantity to form a suitahle fusible slag with the silica. Oxidized arsenic compounds react with iron and carbon producing a speiss, often of the form F^As, and require extra fuel. The reaction is as follows:
The molten products separate at the hearth according to the specific gravity, that of lead being 11.5, speiss 6.0, matte 5.2, and slag 3.6. The lead, collectJed in the crucible, is withdrawn at the lead-well. The alag, matte, and speiss are drawn off at the alag tap-hole on the level of the top of the crucible and of the lead. The separation of nmtt« from slag is generally effected in a Fore-hearth outside the furnace.
Slags In Silver-Lead Smelting
The object of silver-lead smelting is to reduce the lead, and incidentally the gold and silver, from the ore. The sulphiu* present forms, with the copper, iron and a part of the lead, a complex artificial sulphide (matte), while basic flux is added to form a slag of the composition that experience shows necessary.
Slags are silicates of extraordinary complexity; and not all merely fusible slags work well in a ^ver-lead blast furnace. Type slags arc those so proportioned in silica, iron oxide, and lime as to work well in a blast furnace. Slags that vary from the proportions become defective in operation. To fulfill the requirements of good slag, it should have, in the normal operation of the furnace, not more than 0.7 per cent lead, or 0.5 oz.-silver per ton, when producing base-bullion not higher than 300 oz. silver per ton. The density should not be greater than 3.6. It should not permit accretions to form at the hearth, nor the crraping-up or appearance of over-fire. If a alag varies from one of the types given, it is either poorly reduced or makes other trouble in the furnace. Thus a slag of the three-quarter tj'pe, in which the CaO falls to 20 per cent (the other constituents being given in the table), ia found to contain, for example, 1 per cent lead, and more thwi 1 oz. silver per ton; that is, it is " dirty." It easily may happen that a dirty alag is fusible, but we know that the slag will work sat^actorily if correct according to the type, the other conditions of good running being in evidence. Indeed, it ia a cconmon experience that a furnace, worldng poorly on an incorrect slag, begins to run well when a correct sl^ comes down. Following we give a table of type-sla^ that have been foimd to work well in practice :
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Slag In Silveu-Lead Smelting
Table Of Typical Slags
i
SiO,, Psr Cent.
C.tB.,MrtO,
Per Cent.
Qu«teMl« i C
28
12
Half-oUg 1 E
Halt-flUg J
Whole, or 1 to 1 Blag > G
28
According to the ratio of CaO to FeO the slag is called a " quarter," a " half," or a " one-to-one " slag, etc. Thus the sl^ E of the subjoined table is called a half-sl^, the C^aO being but half of the FeO. The slag C is a quarter-^lag, the CaO being quarter of the FeO. In this table of typical well tested Blags the threti elements SiOa, Fe(Mn)0, and Ca{Mg,Ba)0 are calculated to comprise 90 per cent. If the sum varies from this, the ratio ia still to be preserved.
Since any of the slags of the table can be used, the question arises, which is to be chosen? In this we are guided by the economic conditions. If the ore of the district is silicious, and most profit is derived by treating the ore at hand, use a silicious slag thut needs the smaller amount of fiux. If irony or limy ores are plentiful and profitable to smelt, we use them, substituting them for flux. It is found, however, that slf^ of the type , M and G of the table drive more slowly and require more fuel than the basic ones. Perhaps the most satisfactory of the slags, and the one that can be used where silicious ores are plentiful, is the three-quarter slag, F. When a slag of a certain type, for example a silicious one, is not working well and is forming accretions in the furnace, a radical change to & basic type is found beneficial, or from a bade slag to a silicious one.
Action Of Various Bases In Slags
Iron. — Iron ore is quickly reduced to ferrous form under the action of the CO in the furnace or of the highly heated fuel thus:
Iron oxide, being a stronger base than lead oxide, replaces it in the slag, and the latter ia reduced by carbon to metallic lead.
PbSi03-hFe0+C = FeSi0-|-Pb-|-C0.
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Manganese. — ^The equivalent for manganese U 55, and for iron 56, and they are reckoned as having equal values for fluxing. Manganese is found in fiome of the Leadville iron ores to the extent of 10 to 15 per cent, and since by introducing another element, it adds to the complexity of the slag, it also adds to the fusibility.
The Alkaline Earths. — Lime, magnesia, and baryta act in inverse ratio to their atomic we^ts in fluxing olica, hence to obtain the equivalent in lime, the percentage of magnesia is multiplied by 1.4 and of bar>'ta by 0.4. A slag, high in Ume and consequently low in iron Uke the last three in the table, is of low specific gravity. Its use thus results in a better separation of slag from the heavier matte. Lime being a stronger base by one-half than iron, and generally a cheaper flux, the tendency is to chooee the limy slags. It is noticed that the higher the silica content of the slags of the table the higher is the lime, and that high silica calls for high lime. Dolomite, having a high content in magnesia, generally is avoided in silver-lead smelting, for it tends to make stag pasty and streaky, and the imfavorable effect is aggravated when zinc is also present. Two analyses of limestone and of dolomite are given below to show conditions typical of actual practice.
Canyon City Limestone. — CaO,49.8 percent; MgO, 3.0 percent; SiO, 3.1 per cent; Fe, 0.8 per cent.
Iron Coun^, Missouri, Dotomite. — CaO,26.6 per cent; MgO, 17.6 per cent; SiOa, 5.1 per -cent; Fe, 3.3 per cent.
Fluorspai. — ^This has no unfavorable, but rather a favorable effect upon the quality of the slag. The fluorine, however, uses CaO, and hence the slag must analyze higher in CaO than the type requires, or it will not be clean.
Ahunina. — It is uncertain whether alumina acts as an acid or a base. It is sufficient for the purpose of silver-lead smelting to regard it as a neutral constituent that dissolves in slag and acts in neither way.
Zinc. — Either blende or zinc oxide causes difficulties in the blast-furnace, the Uende being the more objectionable. Blende is in part decomposed in the presence of iron to zinc oxide, but the zinc in any form tends to make a stiff, pasty, difficultly fusible slag. It may be regarded, like almnina, as being dissolved in the slag. It goes into both the slag and the matte and diminishing the specific gravity of the latter it causes a less perfect separation of the two. Where much zinc is in the chai^, it is customary to modify the type-slag by calculating the zinc oxide as replacing one-half the percentage of Ume. Take, for example, the half slag J of the table.
In the first colimin we write the slag as the type requires. In the second column we add the 8 per cent Zn and reduce the lime by 4 per cent l^ which the total becomes 94. Since the c<Histituents should amount to but 90
Nature Of Slag As Ajtected By Bases
Per Cfiit,
With
Percent,
Per Cflnt-
31
31
8
FeO
90
94
per cent, all are reduced proportionately in the third column so S8 to give 90 per cent as the sum.
Copper. — Copper present in the chai^ enters the matte when, as generally is the case, sulphur is present with which it can combine. In amelting carbonate or oxidized ores, which furnish no sulphur, the copper becomes reduced, and enters the Irase-bullion, pving a lead so droesy aometimes as to clog the lead-well, and accumulate and solidify in the crucible. The remedy is to supply sulphide to form matte into which the copper can enter.
Antiinooy. — ^Either as an oxide or a sulphide, antimony is reduced like lead. It alloys with the base-bullion, making it hard, and is removed and recovered later in refining the base-bullion.
Arsenic. — This frequently b encountered in silver-lead smelting. When present in small quantity it is volatihzcd, but in large quantity it forms a speias. Where it is intended to produce a speiss, iron is provided with which the arsenic unites. In the fire-assay of arsenic-bearing lead ores, a bead of speiss is foxmd attached to the lead button. Prom the percentage of this we can compute the we^ht of the speiss that will be formed ; and we may assume that 70 per cent of it is Fe, Where a direct determination of arsenic is made we can compute the weight, and multiply this by 2.3 to express the quantity of Fe to be provided on the chai^ for the purpose.
FDEL m SILVER-LEAD SHELTINO
The fuels used in silver-lead smelting are coke, charcoal, or a mixture of the two. Wood and hard coal have been used experimentally, the former in cert^ cases of scarcity of fuel.
Coke. — Coke is the kind of fuel commonly used. The ash viiries from 10 to 22 per cent, and the fixed carbon from 89 to 77 per cent. In coke of high a^, not only is the ash to be smelted, but the carbon is coiv respondingly low, so that the coke is less efEcient. A great difficulty wilii high-ash coke is that it is often friable, making accretions or scaffolds. Analyses of two typical samples of coke give the following results :
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Connellsville coke contains fixed carbon 87.5 per cent, ash 11.3 percent and sulphur 0.7 per cent; El Moro coke, fixed carbon 77.0 per cent, ash 22.0 per cent and sulphur (when the coke is made from unwaged coal) 0,9 per cent.
lu computii^ a chat^ the coke-ash is taken into account, analyses being as follows: Ash of ConneUsville coke: Si02, 44.6 per cent; Fe, 15,9 per cent; CaO, 7.0 per cent; MgO, 1.9 per cent; ash of El Moro coke: Si02, 84.5 per cent and Fe, 5.0 per cent.
Charcoal. — ^This fuel is used in distiicte far from railroads, where the cost of coke is high. It is a good fuel for oxidized ores, but is friable and makes unde^rable fine which may form accr^ons or scafi'olds in the furnace. It renders a char^ more open than coke, and contains less tlian 2 per cent ash. Coke weighs 25 lb. and charcoal 10 lb. per cubic foot when loose, the weight of a bushel of charcoal being 14 to 16 lb. Even where charcoal is cheap it is desirable in operating the furnace to use part coke which, fed to the walls, bums more slowly than charcoal and makes the tUyere-zone hotter and gives a more liquid dag.
Quantity of Fuel. — ^This varies according to the nature of the charge, and generally is from 10 to 15 per cent. Charges that contain sulphur and make matte need less fuel than oxidized ores. Only sufficient is used to give adequate reduction and a hot slag; and the metallurgist is guided by these requirements in adding the fuel.
Pulverized Coal. — This is now being used to supplement the coke fed with the charge, being injected at the tuyeres of the blast-furnace, and burned as it meets the glowing charge. It is being used not only for copper but for silver-lead furnaces.
In smelting ores high in zinc, this, liberated by the fuel and flux, soon coats the descending coke with a white coating of zinc oxide, so that it bums with difficulty; the temperatm^ falls, the slag becomes pasty and works poorly, and accretions form, so it is evident that the more fuel that can be added at the tuyeres the better the fiimace should run . It also suggests the smelting of such ores in the revcrberatory rather than in the blast-furnace.
Calculation Of A Lead Blast-Fusnace Charge
When sulphur-bearing, oxidized, or silicious ore is used, we have to consider not only the sulphur, silica, and other constituents of the ore, but also the products of the furnace that remove the constituents.
Ore (galena for example) containing less than 10 to 12 per cent sulphur generally is smelted without roasting. It is cheaper to do this, for by roasting, the sulphur of the ore is reduced to but 3 to 4 per cent. Many ores within the above limit are leady ores, and difficult to roast because of the
Calculation Of Charge 485
fusible nature, but the matte that they produce is easy to roast for the elimination of eulphur.
Ore intended for roasting may be simple, consisting of iron sulphide, or complex as shown by the following analysis of a roasted ore: SiOs, 10 per cent; Fe and Mn, 27 per cent; CaO, Mgo, and BaO, 2 per cent; Zn, 8.8 per cent; Cu, 0.4 per cent; 8, 6 per cent; Pb, 35 per cent, and Ag, 50 oz. per ton. The base in the roasted ore was present as sulphide in the raw ore.
The so-called oxidized ores coast of the carbonate of lead with a gangue of iron oxide, limestone, dolomite, and ffllica. Such ores though called oxidized, often contain a little sulphur, as sulphide (galena or pyrite) or as sulphate.
Silicious ores are added to charges, in spite of the large excess of ralica, because the gold and silver are present in quantity to pay to recover. The lead of the charge takes the gold and silver contained in such ore, while the silicious gangue is fluxed into a barren slag and sent to waste.
Both iron ore and limestone are added to the charge for fluxing the silica, making a slag of a predetermined composition or type. If the fluxes contain gold or silver the metals can be recovered, since they go into the base-bullion or work-lead. Without gold or silver they are called barren or " dead " flxixes. Ore carrying an excess of iron or lime over silica (called iron or lime excess) is in the same category, since the excess is useful for fluxing, and is credited in purahasing ores. Thus, ores containing 10 per cent SO2 and 40 per cent Fe are said to cany 30 per cent iron excess.
Not all the slag that issues from the furnace is clean. At the spout where the matte Hows, and in the shell lining the cavity of the fore-hearth, slag, containing drops of lead and matte, is found. When a slag-pot is emptied at the edge of the dmnp, there remains a shell or coating of soldi-fled slag. This shell, half an inch thick, is found to contain drops of matte that did not entirely settle in the fore-hearth. This is particularly true when the fore-hearth has formed a thick lining and soon must be replaced by another. All this slag having value, and called "foul slag," is an acceptable addition to the charge because of the fusibility, and the coarse condition, permittitig free passage of the air of the blast.
Computation of the Charge. — ^To determine the amount of the fluxes (iron ore and limestone) to add to the charge, to give a slag of a desired composition, it is necessary to know the weight of the ores to be used, and the results of analysis of the ore, fluxes and fuel, and also the composition of the slag and matte that are to be produced.
When a charge, thus calculated, has been put on the furnace and, after several hours, has come down, that is, has come into the melting zone, BO that the slag of it begins to flow from the furnace, then a sample can be
taken, analyzed and the result of the an^ysis known in two or three hours. If this result shows variation from the desired slag, the cha)^;e-composjtion may be suitably altered to correct it. However, before making any changes, we must note that the slag is hot and weU reduced.
For a charge to contain sintered ore, of size suited to automatic or mechanical charging, the following is an excellent example: Of this chai^ some 70 per cent is sintered material of lead. To this has been added of silicious ores, A and B 350 lb. B^-house fume is a product burned at
Blast-Furnace Charge Ho. 8
Wmoht.
Pb.
SJOi.
C.
w.,.
Dry.
Ct".
u,.
Cent,
Lb.
Per Codt.
Lb.
P«
C-nt
Lb.
Crn'l
3S0
1S§
ino
80D
is
3
sot
s'.f
i'.o
Iit
Isrili,::::;;
Ifi
W«,ht of d>vp>.
F»0+MnO in 8181=1107
Sloa-SiOi -31 permit Total ■!■■ -3130 lb.
JfaUi.Pb-lS.Spenicnt: Fe -*5 per tent; Cu. 7.8 per cent; S -22.5 per cent; Zn-e.O per cvit.
the bag^house into a coherent product. To these is added iron ore and limestone to form a slag to contain SiOa, 31 per cent; FeO {and MnO); 35.5 per cent; CaO, 17.5 per cent. There is 8 per cent or so of zinc on the charge, and the slag composition corresponds most closely to the recalculated, zinc-bearing charge, given under " zinc." The slag used is higher in silica, and not so clean as the type-slag; on the other hand it runs with less furnace trouble. Not coimting the coke or the slag, the chai^ is 4900 lb. in which the fuel is 14.3 per cent. The slag (3130 lb.) carries 1.5 per cent or 40 lb. of sulphur. Of the sulphur 30 per cent or 53 lb. is volatilized, and this leaves 84 lb. to form matte. It will be seen that the ratio of sulphur to iron gives a factor of 2, so that 168 lb. of iron is needed for the matte, leaving 860 lb. of iron (equivalent to 1107 lb. of FeO) to enter the slag. Since the matte contains 22 per cent of sulphur the 84 lb. present in it should give us 380 lb., equal to a matte-fall of 7.7 per cent only.
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Chapter Xxxvii Products Of The Blast-Furnace
. Flde-Ddst
A blast-furnaee, 44 by 164 in., takes at least 6000 cu. ft. of air per minute when in full operation. The escaping gases, of an average temperature of 150° C, have expanded to 8000 cu. ft., and have a velocity, while rising through the charge, of from 5 to 10 ft. per second. When with a closed-top furnace the side doors are opened, additional air iB.drawn-in and particles of 20-mesh size may be earned into the down-take and the long main flue leading to the tall stack which produces the draft. This main flue, which is common to all the blast-fumaces of the plant, is of large cross-section, for the purpose of settling and collecting the particles called " flue-dust." The collected dust, after suitable preparation either by sintering or by briquetting under pressure, can be returned as part of the blast-furnace charge. It commonly amounts to 0.5 to 0.8 per cent by weight of the charge, but with a sinter chai^ it should be even lower than this. These figures refer to what is caught in the flue. Where a bag-house is used, its saving should be added.
Some of the lead and silver and much of the sulphur, dnc, and arsenic of the charge is volatiUzed. This, in part, adheres to the cool surface of the flue. Eventually it flakes off and falls to the bottom, and is there recovered. Flue-dust is therefore composed of (1) duat carried along by the flue and (2) lead fume, containing the other volatile metals, condensed on the cool surface of the flue.
When a bag-house is not used, not all the material is recovered; the finest part may escape. An analysis of flue-dust made at the Pueblo Smelting Works, Colo., shows PbO, 37.6 per cent; ZnO, 53 per cent; FeaOj, 25 per cent; AbOa, 1.3 per cent; CaO (from the limestone), 5.3 percent; Si02, 8.9 per cent; S, 2.5 percent; SOa, 1.6 per cent; HjO, CO2 and C (from the coke), 11,2 per cent. When arsenic is present, part condenses in the flue, but much of it escapes, and is thus happily got rid of.
The carrying power of the moving gases varies as the square of the velocity, or directly as the draft pressure, therefore to settle out as much flue-dust as possible the main flue should have a large sectional area, or better yet, if the bag-house is not u%d, a dust-chamber should be provided.
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488 Products Of The Blast Furnace
Flues have been made of sheet-eteel, but the metal corrodes under the action of the sulphuric acid aiid sulphates that are in the flue-dust, BO that they last about ten years; brick therefore remains the favorite material for such construction. The bottom of brick flues is frequently a series of steel hoppers. Since these are continually covered by the flue-dust, they are protected from the fumes, and last a long time. A flue, rectangular in cross-section, may have brick walls, and top, and a hopper bottom set at such a height as to leave room beneath for a car running on a track at the ground level. The car can be set under any hopper, and to avoid escaping dust, the contents may be drawn into the car through a canvas sleeve fitted over the spout.
The Bag-House
The Bag-house. — Fig, 285 is a transverse section of a bag-house 140 ft. long by 24 ft. wide, as used for the recovery of flue-dust and lead fume. Altering the blas4>-fumace ga.ses or fumes through cot^ ton or woolen " bags " that leave them colorless. It is of a size to take the gases from dx furnaces, equal to 60,000 cu. ft. per minute.
From the main furnace flue the gases are drawn throT^h a 12-ft pipe called " the trail," by a No, 14 suction-fan which delivers them under pressure through a 7-foot pipe to flue A of the bag-house. Just before entering the fan, at the top of the flue, there is fed in a small and regular supply of finely ground quicklime to neu- Pio. 265. — Bag-house. tralize the small amount of
H2S0,i in the gases, which,
if allowed to remain, would speedily corrode the bags.
The bag-house, 140 ft. long, consists of a chamber " B " and a bag
chamber " A," and is divided by transverse partitions into five chamben
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Bag-House
or bays. Each bay has its own inlet and its own exhaust pipe both controlled by disk-valve8, and an exhaust pipe, blanching into the p^ X, which leads to the trail. As shown by the arrows the gases pass by the itUet into the chamber " B " enter the bi^, inflating them, and filtering through the interstices of the fabric, escape through the flue " D " to the main stack (not shown) 210 ft. high. In the course of eight hours the bag^ are getting clogged by adherent flue-dust and this should be shaken off. By adjusting the two disk valves, above mentioned, any bag is by-passed, the bagp of it are collapsed and the flue dust that has accumulated on them is loosened and falb down into the chamber " B " of that bay. The bags arc ^haicen several times at five^mnute intervals to ^ve the dust time to faQ down. Once in several days a side door of the bay is opened and a man entering takes hold of and gives the bags a more thorough shaking, the better ' to clean them. The dust sliding down the steeply inclined bottom of the bay is carried by a helical screw as m Fig. 367 to a discharge opening under the division wall of the bay. The " thimble floor " of the bag chamber " C " is of f-in. sheet steel, pierced with holes for thimbles. Fig. 266, these being 1 1 in. in diameter by 10 in. high, there being 240 per bay, or 1200 in all. The bags arc about 13 in. in diameter and are 42 ft. long, wired at the bottom to the thimbles and, by means of wires, closed and suspended from 2^n. pipes at the top.
Treatment of the Flue^usL — ^The dust contains 60 per cent of lead, and is extremely fine. When 2 ft. in depth or so it is ignited, and once started, combustion proceeds, causing the dust to become sintered together, and in a condition favorable for feeding into the blastr-fumace. The burned dust contains oxides and sulphates of the metals, viz., Pb, 60 to 70 per cent; Zn,3percent; Fe, 0.5 per cent; As, 1.3 percent, and Ag. 4.02 oz. per ton. Since the flue-dust is resmelted and the arsenic volatilizes, it tends to increase in the bag-house product; bismuth accumulates in the same way.
Fio. 266.~Detail of Bag-bouse Thimble.
BRIQUETTIHG FlUE-DUST
Fhie-dust can be wet down and fed back to the blast-furnace. If fed a little at a time, it is simply carried ^ain into the flue, but while wet, in occasional lai^e charges, it may be fed so that most of it is carried down and smelted. The efiective way is to make it into briquettes with milk-<rf-ly
490 Products Of The Blast Furnace
lime B8 a binder. Fig. 267 rcpreeents & plant containing a White briquet-ting-presa for making briquettes composed of flue-Khist and milk-of-lime to which is added fine roasted ore. At the right in the figure, shown to be on a high pbttform, b a pile of quicklime. This is fed, together with water, into the lime-mixer, a trough divided transversely by a partition. One compartment is shown aa containing the lime being mixed to a thin paste, while the other is now empty. The paste is drawn from either compart-
Fio. 268— Horizontal Pug-mill.
ment to a horizontal double-shaft pug-mill. Each shaft is provided with mixing-blades. Flue-dust from the pile at the front of the lower platform is shoveled into the pug-mill and thoroughly mixed with the milk-of-lime by the revolving blades of the pug-mill which, being set at an angle, propel it to the discharge-opening immediately over a troughed conveying-belt. It drops into the hopper of a six-mold briquetting-press where it is made into briquettes that drop upon a flat conveying-belt delivering them to a pile. The briquettes may ]k used at the blast-furnace freshly made, but the usual plan is to dry them, as clay-bricks are dried.
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Metrods Of Matte Treatment 4S1
Fig. 268 illustrates the internal construction, the spiral screw at one end of the shaft acting to speedily discharge the mixed material.
At times the briquetting is omitted, and the pug-mill mixture is wheeled to a drying-floor, or ia distributed evenly upon one of the orB-beds. By the time the bed is used the mixture has set and beocuies a hard masa capable of withstanding handling without being broken.
Lead-Copper Hatte
llie lead-«melting chai^ generally contains copper, and the copper accumulates in the matte. Since matte is roasted and returned to the blast-fvimace, the content in copper gradually increases. When increased to 12 per cent, the copper matte is again roasted and treated in a separate blast-furnace, with silicious ore and oxidized copper ore, to produce a matte of 40 per cent copper, called " shipping^latte " because it often is shipped to a copper works to be treated for copper. This operation is called " concentrating."
Cohparisoh Op Hatte-Treathent Hbthods
It has been ui^^ against the treatment of low^rade matte that from each ton treated in concentrating it there is produced a ton of slag, and half a ton trf lead-zinc fume both of which have to be retreated in the blast-furnace. But in the case ci shipping matte, 3 tons of low-^rade matte having been made into 1 of ^ptnng grade, the consequent slag per ton of matte is but half a ton and the percent^ of lead is proportionately lower.
lie aven^ compoeition c4 shipiHng matte may be thus ^'^'^^
Pb, 26.9 per cent; Cu, 43.1 per cent; Ni and Co, 0.4 per cent; SOs, 0.3 per cent; Fe, 8 per cent; Zn, 2.5 per cent; S, 15.5 per cent; As, 1.7 per cent; Sb, 0.76 per cent; Bhowing the complex nature of such matte, and how it takes up every impurity.
A satisfactory way of treating roatte, where it is willed to produce a more finished product, is to crush it to a 4T-me8h size and to roast it. It is next sent to a blast-furnace and again smelted with alicious and oxidised ores of hi^ grade in copper. There results a matte of 65 per cent Ou and a certain quantity of " bottoms," the result of the sepEtration of copper from the matte. The bottoms are cbai^!;ed into a reverberatoiy furnace through rad&^looks, and the coarse-broken matte is put on top. The doors are closed, and the charge is fiied with an oxidiidng Same, as in the Welsh procesa <rf " roasting." The charge having melted, a reaction of the cuprous oxide on the cuprous sulphide takes place, as described in the Welsh process of making btieter-copper, see page 389, and tiie charge
492 Products Of The Blast Furnace
beccmiea reduced to aji impure copper containii^ aTBenic, bismuth, aad aotimony, as well as the gold and silver that were contained in the jixatt«. The copper is then poled to reduce the cuprous oxide, and ladled into anode-molds. The anodes are sent to an electrolytic copper refinery for treatment.
THE CONVKtTraO OF LEADT MATTS
One smelting works in the United States, the Tooele pluit of the International S. & R. Co., has one department for the smelting of alv-er lead ores, another for the reverberatory smelting and converting of copper ores. It is therefore eaoly possible to treat the matte from the mlvei^ lead furnaces in the converter as follows:
The matte is tapped fnun the blast-furnace foie-hearth into lO-ton pot^ and at the conv^ler department is poured into similar pots standing in a pit. The final portion is held back to be potu^ into a shallow cast-iron pan together with some lead that has separated out in the bottcnn erf the pan. When cold enough the crust of matte is lifted off, leaving the lead, still molten, behind. The main body of the matte in the 10-ton pot is tranderred by crane and poured into the converter. At the beginnii^ of the Uow copious fumes are evolved which are drawn by a suction-fan to a bag-house. Special care must be taken that these fumes do not get too hot so as to aet the bags afire. Pyrometera comiected to signal lights indicate when this may happen and the converter mtm may then turn off the blast until the temperature becomes normal again. The quantity of fume, d both lead and zinc, diminishes as the blow proceeds, the smaller quantity of zinc fume uniting itself to any SO3 present to form zinc sulphate and so preventing acid corrosion of the ha^. The bulk of the fume is lead, being about 65 per cent t^ the totaL This fume, collecting at the bottom oi the lower chamber of the bf^-house is set afire from time to time, forming a slightly dntered product to return to the blast-furnace.
The converter is blown without addition of mlica, thus forming a slag high in magnetite, this magnetite forming a baac lining that is permanent. The slag itself ia basic. The contents of the ladle go to a granulator, where it is poured, the stream of slag being hit by a flat jet of water which - granulates it. The product sinks into a deep hopper filled with water, out of which the slag is raised by a bucket-elevator. Added to the roaster charge, it forms an acceptable item of the enter charge because of its excess of iron.
In a typical instance (per t«m of blister copper produced), there is yielded in the flues and at the bag-house 0.368 ton of fume, yielding 0.221 ton of lead. An analysis of the fume showed Pb, 52.5 per cent; Cu trace; Zn, 3 per cent; S, 5.4 per cent; As, 14,2 per cent; Sb, 1.6 per cent; Fe, trace; Ag, 10 oz. per ton. This Indicates that copper is not
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PRICE OF MATTE 4d3
volatile, while arsenic is quite so. The f\tme, after burning, ia sent to the blast-f umace to recover the lead.
Returning to the matte, now freed from lead and sine, it is added to the copper converter charge, the whole being blown to blister copper.
Selling Price Of Matte
Gold is paid for at S19 per ounce, silver at 95 per cent of the New York quotation, copper price to be that of electrolytic at New York quotation. Lead over 5 per cent is charged for at 30 cents per unit, zinc over 7 per cent at 20 cents per iinit.
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Chapter Xxxviii Production Of Lead Ores And Prices
Costs Of Lead Ores
To illustrate the method of calcukting the actual cost of treating an ore, as in the Colorado or Utah silver-lead smelting practice, we take the case of a so-called neutral ore (Si02 equal to Fe). The ore is assumed to be oxidised, to contain lesa than 5 per cent sulphur, and at least 10 per cent lead. It is to be treated at a works having an output of 400 tons of charge daily.
The cost of treating a ton oi chai^, and of treating a ton of the ot« including the flux, is as follows:
Com p»r Ton.
On.
E)epreeifttion 10 per cent on $2,000,000 investment of plant.. . .
The figures in the second column are obtained by multiplying the total weight of the chai^, 1.4 tons, by the cost of each item per ton of material Mid then adding the cost of the fhix. This corresponds to the ^;ure above obtained, and may be stated ag^ as follows:
1.4 tons of materifJ (1 ton ore, 0.3 ton limeatone, and 0.1 ton of iron ore) at $4.83
p«r ton at charge for smelting 16.86
Coat of fluxes, 0.3 ton of luneetone at $2.00 0.60
CoBt of fluxea, 0.1 ton of iron ore at $8.00 0.80
In case the ore contains sulphur in quantity to require roasting, $2 should be added. For sulphiu' over 5 per cent and up to 10 per cent add
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Ore Pric3E8 495
30 cents per unit to cover the expense of iron ore for diapoeing of the extra sulphur, and roasting the matte made by it.
Distribution. — The costs of production per ton of charge smelted have been thus divided: Labor 23 per cent; coke, 40 per cent; coal, 5 per cent; limestone for coke ash, 6 per cent; maintenance and repairs, 5 per cent; delays due to accidents, strikes, etc., 5 per cent; Sue dust recovray, 2 per cent; administration, 7 per cent.
Obe Prices; Hississifpi Vallet Lead Shelting Woues
N<Hi-a]^entifeTOU8 lead concentrates are bought at a quoted rate based upon an 80 per cent lead content as determined by wet assay, with a deduction of 50 cents per unit for all below and an addition of 50 cents for all over 80 per cent.
Ore Prices; Colorado And Utah Silver-Lead Smelteries
SUver-lead and Dry Ores. — The price per ton, dry we^ht, delivered at the smelting works, depends upon the value of the contents of the ore as determined by fits assay, and based upon the New York price of the metals. From these values must be deducted the charge for treatment or working charge and a reasonable profit. The working charge (W. C.) varies with the lead contents, the insoluble reddue (approximately the silica), the iron, the sulphxu', zinc, and speiss (iron arsenide) present in the ore.
A central custom silver-lead smeltery buys and combines for smelting a variety of ores to its profit; and smelts, not only argentiferous lead ores, but lead-free or dry ones. This it can do if it has enough lead-bearing ore on the charge to insure the extraction of the precious metals from the dry ores also being smelted.
The Metal Values. — ^These are commonly paid for as follows : Gold at $19 per ounce. Silver at 95 per cent of its New York value less a further deduction in Utah of 3.5 cents per ounce to cover the freight charge now made upon base-bullion that contains it, but no such charge is made upon silver' in Colorado ores. The 5 per cent deduction made in the silver price is intended to cover that lost in smelting.
Lead. — ^Based upon a quotation of 6 cents per pound a deduction of 10 per cent is made to cover the smelting loss and in Utah a further deduction of S35 per ton of lead to cover freight and refining loss. This is based upon a chai^ of tl7 per ton of lead for freight and a further deduction of $18 per ton for re&iing.
Treatment Rate or Working Charge.— The base-price for smeltii^ ores is $2.50 per ton with lead at 6 cents per pound; if the value of lead exceeds this, then 90 cents per imit is added to the treatment rate, while, when
'Ms Production Op Lead Ores And Prices
lead falls below 6 cents, a corresponding 50 cents a tniit is takeD trom that Tat«. No distinctioB is now made between oxidized and sulphide ores. At present the furnace charge is two-thirds or more sinter, and in this has been put the fines out of all oiee, such fines coming from the custom of crushing when sampling. Since the ooet per ton of ore fumaced, and the cost per ton of materials put through is accurately known, as well as the cost of roasting and refining, it has been thought best to make the base price for ranelting equal to this actual cost phis a reasonable profit and to seldom make finer distinctions. This runs much the same day by day, and Ha month's profit, in normal operation, approaches the known figure. When lead ores are scarce the ore buyer will make concessions to pn>cure them. In Colorado a different rule holds and in the schedule for dry oxidized ores the treatment charges vary from S7.50 to $9.50 per ton between gross values of from $8 to S20, that is t^iese charges increase with the value of the ore.
Debits or Penalties. — A chaige is made of 10 cents per imit for all insoluble matter. Speiss over 5 per cent is charged at 20 cents per unit. All zinc over 10 per cent is charged at 30 cents per unit. All sulphw is charged at 25 cents a unit, but not to exceed a maximum of $3 per too.
Profits Per Tor
We will now compute the price f . o. b. at the works for a Utah ore of the composition, Insol., 40 per cent; Fe, 10 per cent; CaO, 5 per cent; Pb, 20 per cent; Cu, 0.3 per cent; Zn, 8 per cent; S, 5 per cent; Ag, 18 oi. and Au 0.10 oz. per ton.
Credits.
MeUl valuai Au 0.10 oi. at S19 SI -00
Ag 18 oi. at 11.10 per oi. 95 per cent of this leas 31 cento 18. 18
Pb 20 per cent or 400 lb. lees 10 per cent leee 1.75 cents with lead st
Sit 6 cents per pound 15.30
Metal VBluea S35.38
Pe credit 10 per cent at 6 cents 60
Total credito S34 . 78
Debits.
Insoluble 40 per cent at 10 cents 4.00
Zinc 5 per cent over allowance at 30 cents 1 . 50
Total debits $8.00 98.00
Net value of the diy ore S26.78
The realization per ton will be
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Smelting Costs 497
Gold, lOOpercentof 0.10 OB. at $20,56 perounce S2.0e'
Silver, 98 per cent of 18 oi. ftt $1.10 per ounce 19.40
Lead, 92 per cent of 400 lb. St 6 oenta per pound 22.08
Total metal values $43.54
Net ooet per ton f.o.b. fit the works 26.78
Treatment 6.00
Freight at $17, refining at $12 on 0.184 ton base bullion 6.34
Selling the product 0.33
Interest on metals in process, three months at 6 percent 0.S3
Total costs $38.97
Profits per ton to pay for capital investment to balance 4.57
Total realisation $43.54
Since analysis is based on the dry ton the moisture must be deducted from the gross weight to determine the weight dry. No chai^ is made, for sampling except in lots of a few tons.
Variation In Costs D0E To Ootpdt, Etc.
Comments on Costs. — ^Where operations proceed emoothly where slag losses are no greater than given for good smelting, where by the bag-house and the electrostatic treater the metal losses are low, where gold and copper in small amomite are not paid for, and where the metalfi are skitlfully sold, profits may be as high as above calculated. On the other hand, if iron ore and limestone must be purehased for fluxing, if the metal market is a falling one, if it is not possible to get the best combination of ores for profitable smelting, if the works are running at part capacity, these profits will dwindle. The money to carry a stock of ores -and supplies, and for freight advances should be valued at 6 per cent per annum.
To treat a ailicious ore of 50 per cent silica, provided the iron ore and limestone must be purehased, is expensive, since with 1 ton of ore one must use 1.5 tons of fiuxes and smelt 2.5 tons of materials of the chai^, the estimated cost being (21 per ton. The actual chai^ for treatment is about half this and even the extra charge of 3J cents per ounce of the silver does not compensate. In general there is an excess of iron for fluxing which comes from iron sulphides and irony ores. It must also be remembered that a ton of such ailicious ore will produce 1.75 tons of slag, which sli^ carries off Mlver and lead. The term " displacement " refers to that condition where much silicious ore is mnelted, so that for 100 tons smelted 40 tons would be ore, while if the ore is neutral of lOO tons of charge 72.5 tons would carry the profit.
The price may be modified according to the needs of the works. Thus, if lead ores are much needed they may be bought even at a loss while to compensate the more plentiful ores may be bought at a low price.
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Chapter Xxxix
Refining Of Lead And Base-Boluon
Primary lead, or that produced by Binelteries, may be divided into three kinds on the market, viz.: Soft lead, which comes from non-argentiferoira ores; desilverized lead produced from base-bullion; aDtimonial lead, a by-product of the Farkee process. Soft lead from the ore-hearth ifi commonly remelted and poled to remove impurities. Desilverized lead may be divided into common lead suited to making pipe, sheet lead, shot, and lead alloys. A softer grade is called corroding lead for making white lead. Antimonial lead as a base for type metal and bearing-metal con-tuns 15 to 20 per cent antimony.
REFINinG BASE-BULLION
San^iliiig and Handling. — The practice at large silver-lead Bmelting works is now to remett all base-bullion from the blas1>-fumace. Sometimes the lead Is taken in molten condition to the remelting kettle frcsn the blast-fumace. When melted in the remelting kettle it is carefully skimmed, and as in lead refining, the cleaned lead is molded into bars. The humming or dross, containing copper and other impurity, is returned to the blastr-fumace. The copper there enters the matte and the lead again goes to the base-bullion. While the lead is being molded samples are taken from the kettle at intervals, and from the samples the assay-results are obtained. The bars for a 40-ton carload, 800 in number, are stamped with the number of the lot, and are carefidly weighed, twenty at a time. Careful assays are made of each lot, both by the shipper and by the refiner.
At smaller plants the punch sample is taken aa described io the chapter on sampling. The results are exact.
CharacteristicB of Base-bullion. — Lead containing silver, commonly called base-bullion, is refined by the Pattinson or by the Parkes process. Commonly the Paj-kes process is used. The object in either process is to effect the separation of the silver and gold from the lead.
To get a clear idea of the principles of refining base-bullion (or work-lead as it is called in Europe) we first must know the composition. An especially base quahty is represented 1^ the following analysis:
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The Lead Refinery
Lead
Impurities: Cu 0.82
Aa 0.38
Sb 0.71
Fe 0.02
S 0,14
It is seen that base-bullion is principally lead. The problem is to soften the lead by removing the impurities, and then to separate the gold and ^ver from the puiified or softened lead. In studying the process the student should refer to Figs. 280 and 269.
The Sefikert
Id Fig. 269 we have a cross-eeotion of the refinery building showing the course of the base-buUion through it as further elucidated in the flow-sheet Fig. 270.
The bullion, in bars or ingots of 100 lb., pass by an inclined elevator to the softening furnace S, whence the softened metal goes to the zincii^ kettle K. The silver and gold are here removed and the lead yet containing some zinc passes on to another reverberatoiy furnace, where t^
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Refining Of Lead And Basg-Buluom
sine is burned off and the residual lead, now soft, is molded into ban or ingots (see Fig. 276), for the market.
Fig. 270 is a flow-sheet of the process from the receiving (rf the base bullion to the production of the market lead and mlver-gold or dor6 ban, the latter to be parted for the production of gold and silver.
Softbhing Basb-Bdllion
The Softening Funutce. — Softening is performed in a wateHacketed reverberatory furnace, Fig. 271. The rectangular hearth (rf the furnace, 7 by 14 ft. in aiee, is surrounded by a sheet-steel double water-jacket, ediown in eection at (o) in the sectional elevation (c). The jacket aasiEls
o Biut Fanwce Skin to BbutFunMoe
Fia. 270. — Lead-refitiing by the Parkes Proccos.
in reasting the action of the molten litharge formed from the lead in th? operation. Within the water-jackets the hearth and walls are a dense aluminous brick, and at the slag line, where the corrosive action of the litharge b intense, bauxite brick of 98 per cent AI3O3 are used. The furnace is heated by the firebox, having a grate 4 by 5 ft. in dimensions, so that a high temperature can be attained in the furnace. The letters c, c indicate the rear working doors and b, b, b the front doors in which the base-bullion is charged. At the front of the furnace the tap-hole e is provided, through which the lead is tapped when the chai^ is finished. The furnace communicates to a stack 50 ft. h^ by a flue at the front end.
Operation. — The work is done in two st^es. In the first stage at a low temperatxire, the copper is removed. In the second, at a h^h temperature, the arsenic and antimony are expelled, after which there is left only the Bcrftened lead containing the pilous metals.
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Softening Furnace
Soi
The baee-bulUon, in charges ot 30 tons, is placed in the furnace by means of a long-handled paddle or " peel " having a blade 2 ft, long by 6 in. wide. The bars are laid one at a time upon this and placed as deared in the furnace, being piled in a heap on the hearth. The doors are closed and the bare are gradually melted down, the droas contained in the bulUon rising to the top. When melted the heat is maintained sh^tly above Ihe melting-point, but not higher. In about two hours the dross that has risen to the top is carefully skimmed, by means of a long-handled perforated
stdnuner, and removed through the door to a wheelbarrow placed for it. The dross, residue, or skimming, called the " copper skim," consists of a drossy lead containing the iron, sulphur, and (especially important) most of the copper of the base-bullion. The removal of these completes the first stage of the process.* The liquated dross thxis skimmed, which * It should be here noted that where the emelting plant and refinery are in one, the bla8t--funiace lead is treated as described under " Sampling and Handling." There reeidts a product which, when charged into the softening furnace, can omit the first stage and the lead ia Ireated to the second stage or obtaining the " antimony akim "
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fi02 REFINING OF LEAD AND BASE-BULUON
may amoust to 5 per cent ot the charge or 1.5 tons, consists of Pb, 62.4 per cent; Cu, 17.97 per cent; Ag, 0.17 per cent (49 oz. per ton); As, 2.32 per cent; 8b, 0.98 per cent; Fe, 0.43 per cent; S, 4 per cent; and O, 1.87 per cent. Slag, ash, and hearth material also ate contained and must be reckoned in.
The heat of the molten bath is now raised to a bright red (600° to 650° C.) and the flame, made oxidiziog by the admission of an excess of ^ through the thin fire, sweeps over the siirface. Litharge forms, and the antimony and arsenic oxidize and enter the litharge slag. The litharge
Fig. 272.— Howard Mixer.
at this temperature has a corrosive action upon the brick lining, hence the need of a water-jacketed furnace. This sta^^e of the process lasts twelve hours, until a sample of the lead taken from the furnace and placed in a mold and skimmed, shows by the appearance that it is free from arsenic and antimony. Before the antimony is removed the surface of the molten lead will " work," or show oily drops moving upon it. A similar phenomenon is seen in the iirst st^e of cupelling base-bullion high in antimony and arsenic. As the softening proceeds the drops become fewer and smaller, and finally a coating is seen to dull the surface of the hot molten lead, indicating the completion of the softening. For impure base-bullion this stage is of more than twelve hours' duration, and the thick layer of litharge formed retards further oxidation. It is best then to draw the fin
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The Parkes Process 503
and to cool the chai^, to allow the litharge alag on the top to solidify above the liquid lead beneath. The slag is skinned with a long-handled perforated skimmer (compare with Fig. 275), and the charge is fired again if necessary until the impurities are removed. The " antimony skim " consists of the antimonate and arsenate of the lead with a large proportion of lithai^. It is in fact an impure litharge containing 15 to 20 per cent antimony.
The softened lead to be treated by the Pattiuson (see page 510), or the Farkes process for the removal of the contained gold and silver, is now tapped into the desilverizing kettle, 8 ft. diameter and capable of holding 30 tons of lead.
The Parkbs Process
Operation. — The softened lead from the softening-furnace is tapped into a hemispherical cast-iron kettle, shown In Fig. 272, which holds 30 tons or more of lead or the full chat^ from the softener. The kettle is set in brickwork, and is heated from a firebox below. In modem practice kettles aie made large and are 10 ft. diameter and 2 ft. 10 in. deep, holding 60 to 65 tons.
The lead from the softening-furnace Sows along a cast-iron trough to the kettle. In so doing a lithai^ dross, called " kettle dross," forms, and collect on the surface of the metal, and is skimmed off.
The principle of the separation of silver from lead depends on the affinity of silver for zinc, which is greater than for lead. Upon adding and thorou^y mixing in a small amount of zinc it takes up most of the silver. Zinc has a greater affinity than lead, not only for diver, but for gold and copper. When the molten bath is allowed to stand a while, the zinc, being lighter, separates and rises to the surface. At a temperature below the meltjng-point of zinc, but above that of lead, a crust forms that can be skimmed off. Thxis the silver is concentrated in a small bulk of metal, and is later separated from the rich metal by fmther treatment.
The molten bath is heated to an incipient red heat, well above the melting-point of zinc, and cakes or
ingots of spelter equal to about 1.2 p,o. 273._Skimming Base-bullion, per cent of the weight, or 720 lb.,
are added. The quantity required varies with the richness of the base-bullion in silver. The added rinc quickly melts.
Desilverizing machinery b now much used. The most approved
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SSFININO OF LEAD AND BAS&BtJLLION
machine is the Howard, used both for mixiiig and ulriTnTnmg. Fig. 272 lepiesents, in section and elevation, the kettle and the apparatus used for intimately mixing the molten zinc with ibe lead. The machine is brou^t to the kettle by an overhead crawl A and is lowered into it by a chain-block hoist. When lowered into position as shown in Fig. 272, the screw pn>-
FiQ. 274.— Howard Presa.
peller b is set in motion by a steam-driven mechanism so as to produce a downward flow of molten lead in the sheet-iron cylinder a. The cylinder has neither top nor bottwn, and being submei^ed in the lead, a drculation is started, the lead flowing in over the top of the cylinder. Thus a thor^ ough mixing of the content of the kettle is assured. In a few minutes
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Farkes Process 605
the en^e is reveraed, and the flow is made upward over the edge of the cyUnder, then downward to the bottom. The mixing is continued about eleven minutes, after which time the stiiring apparatus is bodily hoisted and moved to one aide. Several kettles can be thus served by one mixer.
The content of the kettle is now allowed to cool two hours or more. The light zinc rises to the top and carries the silver, gold, and copper with it. Finally when the temperature falls below the melting-point, a half-fused, mushy crust or layer forms upon the lead. The crust consists of 65 per cent Pb; 10 per cent Ag and Au; 3 per cent Cu; and 22 to 24 per cent Zn.
Fig. 274 is an elevation of the Howard press by which the sine crust is removed from the lead. Another elevation shows also a sectitm of the castrjron pot into which the press is about to be lowered. In principle the machine is like a cheese-preas. The apparatus is lowered into the lead until the top edge of the cylinder a is but sli^tly above the surface. The phmger or follower c is raised, and the zinc-crust, as it is skimmed frcon the surface, is put in it by means of the perforated skimmer. Fig. 275.
Fig. 275.-
The press thus expedites the skimming. While one man is skimming and puttii^ the skimming into the press, another assists by pushing the crust to one place with a wooden rabble. When full, the press is raised and the surplus lead begins to run out of the half-inch holes in the hinged bottom b. The plimger c is Intnight down, squeezing out more of the lead, and leaves the remaining, mu^y, half-duid mass nearly free from lead, of the composition ^ven above. The press is now run to one mde over a floor paved wiUi cast^ron plates, the hinged bottom 6 is dropped by leleaaing the catch, and the zinc-crust is pushed out by continuing ttu downward movement of the phmger. The crust falb upon the castHion floor-plate, and while soft is readily broken with hammers into hunpe the aze of the fist. Meanwhile the hinged bottom is closed, and the press returned to the kettle and is opened to receive more skimming. These operations continue until the surface of the lead is well skimmed and take in all about twelve hours. The crust amoimts to 3000 lb. and contains 90 per cent of the silver originally in the softened base-buUion, resulting in a concentration of twenty into one.
Dezincing Furnace. — ^Thra fii^ " zincing " removes all the gold and copper for whjch zing has a great affinity. It does not i«nove all the olva, ^
506 Refining Of Lead And Base-Bullion
and the operation must be repeated once of twice more before the Bilvor content is diminished to the fraction of an ounce per ton beyond which it does not pay to go. Of the 1.8 per cent zinc needed, first is added ] of the zinc or 1.2 per cent, then }, or 0.45 per cent, and finally the remaining A, or 0.15 per cent; or 900, 270, and 90 lb. respectively.
The deeilverizcd lead remaining in the kettle after the last sldnunir.g retains 0.6 to 0.7 per cent zinc and traces of arsenic and antimtmy, all of which must be removed before the lead is suitable for market. This ia done by siphoning or tapping the metal from the kettle into a rcverberatory furnace similar in construction to the softening furnace. Here the charts is brought up to a bright-red heat, the zinc is volatilized and burned oR, and Utharge forms as a slag upon the surface of the lead. The operation
Fio. 276.— Molding Market Lead.
takes six hours, and is complete when the zinc has been expelled, as shown by taking a sample of lead in a mold and observing the appearance of the surface as the metal solidifies. The furnace is allowed to cool until the litharge-slag is solid and can be skinned.
Molding. — Finally the lead is tapped into a market-kettle similar to the desilverizing kettle. This is the reservoir from which it is drawn to be cast into molds. The molds, fifty in number, standing in a semicircle as shown in Fig. 276, hold 100 Ih. of lead each, and are conveniently mounted on two wheels by which, when full and cooi, they arc transferred to the adjoining floor. There the lead is tilted out and the molds at once returned to the semicircle to be used again. The lead is withdrawn from the kettle by means of a siphon. It descends into a small cast-iron pot, into which is screwed the 2i-in. pipe that delivers it to the fifty molds, the pipe being quickly moved from mold to mold as filled, without interrupting the flow.
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Retorting The Rich Lead , 507
At the eaid of each round the flow is interrupted onJy to cany back the end of the pipe to the first mold of the series, which meanwhite has been emptied and replaced. The 100-lb. pigs, or bars, are the dedlveriised lead of corn- Dry steam may be blown into the molten lead in the kettle to refine it. It is introduced by means of a pipe inserted deep beneath the surface. The constant agitation produced by the steam brin)^ the metal in contact with the air and oxidizes it and the remaining impurity. It is softer than ordinary desilverized lead, and is easily corroded by the acetic acid used in making white lead. It is accordingly called " corroding lead."
Treatment Of The Rich Lead
Retorting. — Referring to the diagram (F%. 270), we see what becomes of the crust or skimming that results from the first zincing. The material is in lumps, containing 22 to 24 per cent zinc. It is charged with charcoal breeze, into bottle-shaped retorts /, Fig. 277, each holding 1200 lb. zinc-crust. The figure represents at (a) a sectional elevation through the retort, at (6) a transverse section, and at (c) an elevation of a Faber du Faur tilting retort-furnace. The retort rests upon a narrow arch, and carries a grate upon which rests a coke fire that fills the furnace and co-\;ers the retort /. The products of combustion escape by an outlet-port at the back to a stack of good draft. The coke is fed through a hole in the roof of the fiunace, and is poked down, and kept in vigorous combustion, so that a yellow heat (1000° C.) is attained. A condenser, made by cutting off the end of an old retort (as shown at (c), Fig. 277), collects the zinc vapors distilling from the chai^, the condensed zinc being drawn into mold sx through a 1-in. hole, bored through the bottom edge of the condenser. When distillation is complete, the condenser and the supporting truck are removed, the furnace is tilted or revolved by means of a lever on the trunnions, and the remaining " rich-lead " is poured into molds like those used for molding market-lead. The rich lead still retains zinc, copper, and impurities, taken into the crust at the first zincing.
Cupelling. — To obtain the silver (and gold) from the alloy, the English cupelling-fumace, Fig. 278, is used. The principle of the action is much like that of cupellation in assaying, except the litharge here not only saturates the cupel, but flows from it as fast as formed. Fig. 278 shows the firebox b where a long-flaming coal is burned, the products of combustion passing to the chimney. The flame plays over the hearth called a " test," a lai^ cupel rammed tight within a test ring, shown mounted on the carriage. The test is lowered by the jack-screws, remcfved on the carriage, and another is" put into the bottom opening of the hearth.
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Refining Of Lead And Base-Buluon
FiQ. 277.— Faber du Faur Retort.
Fia. 278 — Penpectire View of ED^ifib CupelliDg Furnacea.
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Cupelling
50d
when the first is consumed. The test is hollowed like a cupel, to hold a shallow bath of molten lead 3 in. deep. Fig. 279 shows two views of the test and the supporting truck, including a view of an inverted truck and test. In the furnace, Fig. 278, is seen the overhead pipe, branching to the ash-pit, to supply under-grate blast, and to an opening at the back of the furnace where a tuyere is inserted, by which s stream of air is brought to play upon the surface of the molten red-hot bath of rich lead. The air oxidises the lead to litharge. Other impurities are oxidized and enter the Utharge-elag and are carried away with it. The molten litharge, aa it forms, escapes by a shallow groove or channel in the top of the front edge of the cupel or test. A door w can be lifted to in^)ect the operation, or to cut the channel as needed. At the rear are provided two porte of a sise to permit inserting two bare of rich lead that
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T 1 t r
u« pushed in as fast as the cupellation proceeds. The ends of the bars melt and supply the lead. The Utbarge stream is the size of a lead-pencil, and falls into a small slag-pot beneath. The lead is fed at the rate of 1 to 2 tons daily until the bath has become rich in silver, when the feeding of the lead must be stopped. Oxidation then is continued, cutting the channel deep to allow the remaming litharge to flow out, and finally the mirroi^ like bath of silver appears. The fire must keep the temperature above the melting point of the silver. At the last, a shovelful of bone-ash is thrown on the bath to absorb the remaining trace of litharge as it forms aa the surface. This is skimmed, and the silver is then ready to be ladled out or tapped, commonly Into the casfr^ron molds, each holding 1000 oz. silver. This is then subjected to the acid-parting operation to be described later. The coppeivakimmii^, which is the first obtained from the softening furnace, is returned to the blast-furnace where the sulphiu* of the chai^ combines with the copper and removes it as matte. The reet of the skimming, mostly lead, containing silver and gold, is reduced to base-bullion.
510 Refining Of Lead And Base-Buluon
The third skimniing of the aoftening-fumace, if any, is retunied to Uie blast-fiinisce, since it ctmtains but httle antimony.
The second Boftening akinuning or antimony-skim, containing 15 to 25 per cent antimony, goes to a small reverberatory furnace, 8 1^ 12 ft. hearth dimensions and 10 in. deep, built like a softening-furnace and called a precipitating-fumace. Here it is melted, with a reducing flame into a slag. Charcoal is added, and stirred, to reduce or precipitate part of the lead <rf the slag. The lead, falling to the bottom of the bath, carries - down the mlver of the slag. When the reaction is complete, the supernatant slag is tapped into sl^-pots, and the lead is tapped into a kettle at a lover level and molded into bars. This precipitated lead-bullion is returned to the Boftening-fumace to be softened and desilverized. The antimonial slag, containing about 8 oz. silver per ton, when accumulated, is smelted in a small blastr-fumace to reduce it to antimonial lead of 20 per cent Sb, which is sold to the type founders. The dag is rejected.
THB PATTIIfSON PROCESS
When a kettle containing molten lead is allowed to cool slowly as it approaches solidification, crystals of lead low in silver separate. The metal that remains liquid contains the larger part of the precious metal. The crystals are removed with a perforated ladle, melted in another kettle, Mid allowed to cool. Once more crystals separate that are low in silver, the mother Uquor becoming high in fdlver. If the liquid portion first mentioned be transferred to a kettle and likewise heated and then allowed to cool, the same segregation of the alver into the liquid part continues. We can accordingly arrange a series of kettles containing, at one end low-grade lead, and at the other high-grade, all from one product. A series of this kind, as illustrated by practice at Eureka, Nev., gave the asaya quoted in the table below.
Another crystalUzation would reduce the silver of the market lead to half the value given. The rich lead could be directly cupelled in an English cupelling-fumace, or better, treated by the Parkes process to get rich silver-zinc crust for retorting and cupelling. The process has been modified recently by Tredennick, who raises each kettle by hydraulic power above the adjoining one so that the mother liquor drains from one kettle to the next through & strainer, the lead being cooled near the solidification temperature by introducing steam upon the surface. The cost of operating has been greatly decreased in this way. The chief advantage of the Pattison process over the Parkes is that it gives a product fr«e from bismuth. In the Parkes process the bismuth follows the lead. Bismuth is injurious in lead that is to be corroded to make whit« lead, and it may be necessary to employ the Pattinson process for making a corroding lead from bismuth-bearing ores.
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Bett8 Process 511
Muktt Lud Kattla Ounna Af per Ton. No. 1 1.25
No, 6 18.0
No. e 30.0
No. 7 50.0
No. 8 75.0
No. 9 100, 0
No. 10 150.0
No, 11' 460.0
COST OF RSnHIHG BASB-BOLLIQN
The actual coat of refining baee-bullion is as follows:
Prime or flat-cost, of softeniiig and refining S5,00 to S6.00
General expense 3.00 to 3,00
Lose in metals and incidental expeuaes 1.70 to 3,00
Total SO.TOto J12,00
SELLING PRICE OF BASE-BULUOIf
Base-buUk>n. — A email works may sell this product to an Eastern refinery as follows:
Gold is paid for at J20.40 per oimce. Silver at 99 per cent of New York quotation, sixty days after date of sampling at the refinery. Lead is paid for at 99 per cent of the New York quotation, thirty days after same date of sampling.
Refining cost, 112 per ton. An advance will be ^ven up to 90 per cent of the net value with a charge of 6 per cent for said advance.
THE BBTTS PROCESS FOR THE ELECTROLTTIC REFfHING OF LEAD
The principle of this process depends upon the aolubihty of lead in an acid solution of lead fluosilicate, which is used as an electrolyte. The solution is fonned by diluting hydrofluoric acid containii^ 35 per cent HF with an equal volume of water and saturatii^ with powdered quarts according to the reaction:
In iht) hydrofluosilic acid lead is dissolved. The solution contains 7 to 10 per cent lead, and 8 to 12 per cent of fluosilicic acid (H2S ; Fs) , the free acid varying from 3 to 5 per cent lead. To obtain a solid deposit on the cathode ghie is added to the extent of 0.1 per cent.
512 Refining Of Lead And Base-Bullion
The anodes are plates of the base-bullion to be refined, cast 1} in. thick, raeemblii^ ordinary copper anodes.
The catiiode-sheets that receive the deposited lead are " strip^Hng l^tes," obtained as in the case with copper cathodes. They are made by depositing lead upon steel tathode^plates, prepared for use by cleaning, coating with copper, lightly lead-plating them in the taulis, and greasing with paraffin. On them is deposited the lead, and when the coating is c^ tlie desired thickness the steel cathodes are removed from the bath, and tiie lead coating or sheeta are stripped off for use as cathode. Another method consists in casting the cathodes in the form of thin sheets.
The anodes and cathodes are placed 1} in. apart in the tank. Aa in copper refining, the anodes are in multiple, and the tanks in series. The current enters the anodes, passes through the electrolyte to the cathodes, dissolves the lead from the anodes and deposits it upon the cathodes.
The fall of potential between anode and cathode is 0.45 volt, and the current strength is 15 to 18 amperes per square foot. One ampere depodts 3} oz. lead per twenty-four hours following the ratio of the atomic wei^ts of copper and lead which is 63.6 to 207.
In the process the impuiitiea renuun as an adherent coating on the anode, and consist of the copper, bismuth, arsenic, gold, and silver. The zinc, iron, cobalt, and nickel dissolve in the electrolyte.
The cathode (containing the starting sheet) is melted and cast into bars; the anode mud recovered from the anode and collected at the bottom of the tank is refined to recover the precious metals. When about tbieefourtha corroded the anode is removed.
As compared with ordinary refined lead, electrolyticatly refined lead is pure, being practically free from bismuth, even when much is present in the bsse-bullion, and it must be remembered that bismuth is harmful to " conwling lead."
The residue or anode-slime, averaging 8000 oz. or more of silver and gold per ton, is treated by boiling it with sulphuric acid, usii^ a steam pipe iitserted in the solution to boil uid Estate it with free access of air. The washed re»due is melted in a small basic-lined reverberatory ^unace, the copper is removed by using niter as a flux, and the antimony by the addition of soda. The dati bars finally obtained are parted in the lisual way with sulphmic acid.
The process is used at East Chicago and Trail, 6. C. At Omaha it is confined to the refining of base bullion containing 1.5 per cent bismuth.
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Silver-Lead Smelting Works
SUBLTERT AND REFINERY FOR SILVE|t-LEAD ORBS
We show in Fig. 280 s general view of a company plant for the treatment of their own ores and base-bullion.
Referring to Fig. 280 a double track at the right (in this case the West
side) brings in coke and ore, the first to tiie coke storage and the second to the ore storage bins at the extreme left. Here the coarse ore is taken out to be sent to the blast-furnace building, and the fines go to the Wedge
514 Rehning Of L£Ad And Base-Bullion
roasters and the Dwight-Lloyd sinter plant. The base-bullion frcnn Ibe blast-furnace adjoining goes to the lead refinery, giving two products, Uw market lead to the lead storage warehouse and dor^ bars to be parted at the gold and diver refinery. The blast-furnace flue, parallel to the blast-furnace buildii^, turns at right an^es, forming the " trail " to the fan-house where the fume is driven through the b^-house and finally to tix l&-foot by 200-foot Custodia stack. Powdered coal is used in the refinety and the Dwight-Lloyd mnter plant.
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PART vn
Zinc
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Chapter Xl
ZmC AND ITS ORES
Properties Op Zinc
Zinc is a bluish-white metal of specific gravity 6.9 to 7.2, according to the way it has been cast and cooled. The rolled metal bas a specific gravity of about 7.25. Zinc melts at 419° C. and boils at 950° C. with a characteristic brilliant bluish-green flame. The commercial metal becomes malleable and ductile if heated to 100° to 150" C. and when cooled from this can be rolled. At 205° C. it again becomes brittle, and may be pulverized in an iron mortar. Zmc has a tensile strength of but 18,700 to 22,200 lb. per sq. in. when in sheets or wire. When it passes from the cold solid to the molten condition it increases in volume and on again cooling contracts but slightly. Carbon dioxide readily oxidizes zinc vapor with the production of carbon monoxide and zinc oxide.
To the metallui^ist its value in the recovery of the precious metals in the Parkes process and its usefulness in cyanidation largely appeals.
Zinc ORBS
The principal ' ores of zinc are blende and calamine. In New Jeisey occxu' deposits of franklinite, ZnOFe203. The zinc minerals seldom occur pure; besides the earthy gangue, and sulphides of iron, lead and copper, blende as marmatite or black-jack contains iron so combined chemically that it cannot be separated by orendreeaing methods.
Blende, or sphalerite (ZnS), when of a yellow color as in the ore of the Joplin district in Missouri, is called rosin-blende. When dark in color, due to chemically contained iron as in the ore of the Rocky Mountain States, it is called black-jack. It is from blende concentrate that most of the spelter of commerce is extracted. It needs roasting before it can be retorted or smelted for extracting the zinc.
Calamine is a term applied commercially both to the carbonate (smithsonite) and to the hydrous silicate of zinc. It is an oxidized or sulphur-free ore that needs no preliminary roasting before smelting. On being heated in the retort, the CO2 of the carbonate is expelled, leaving zinc oxide.
Willvnite, the anhydrous siUcate occurring with franklinite in New
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Jersey, mixed with coal, is decompoeed at the high temperature of the retort, yielding one.
It is generally found advantageous to calcine calamine for the purpose of driving off CO3 and water, which are undesiral^ in retorting because of their oxidizing action on zinc-vapor. However, the preliminary caldnation is often omitted, but when perfonned, it is done in kihis much like those in which lime is burned.
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Chapter Xu
ROASTING ZmC ORES
RKDDCnon OF ORES OF ZINC
In outline, the metallurgy of sine conflietB in grinding the ore (generally blende) and n»sting to convert into sine oxide, then charging Uu roasted ore, intimately mixed with fine coal, into horiaoutal, cylindrical, clay letorte, heated to a white heat, where the zinc, reduced by the coal, volatilizes, and the vapor, entering the cool, tapering, clay extension of the retort (called the condenser), condenses there. As it accumulates it ia tapped into a ladle from time to time, skimmed, and cast in molds. When distillation is complete the condenser is removed and the content of the retort taken out and generally thrown away. The cycle (rf operations takes twenty-four hours.
KOAsmiG BUMDB
The aim is to dead-roast the ore, generally to 1 per cent sulphur or less. For every 1 per cent sulphur remaining in the roasted ore, 2 per cent zinc is held back in the retort-residue after distillation.
The ore is ground to about O-mesh size, then lowly and carefully roasted with frequent stirring, finishing the roast at a hi^ temperature to decompose the zinc sulphate formed at the lower temperature. The ore is generally in the form of concentrate, still containing a little gai^ue, galena, and pyrite. To r^nove the final 1 per cent of sulphtu* would require a long time and would not be commercially profitable. The ore is accordingly considered to be finished when it contains no more than that amount of sulphur.
Cheiostrt Of Roastibg Zehc Orbs
We have, in the roasting of blende, the following reactions:
2X43,000 86,400 230,000 71,000- +301,400 cal.
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Thus in an oxidizing flame, blende is roasted to oxide and sulphate, both reactions being exothermic. As indicated in the reactions given in the chapter on Roasting, pyrite or chalcopyrite asmsta in the reactions. At a cherry'-Ted heat the zinc sulfate is decomposed into basic sulphate (3ZnO, ZnS04) thus:
The baac sulphate, exposed to a bright-red heat for a time, reacts thus:
Finally zinc oxide is obtained and the sulphuric anhydride is eliminated.
When limestone or calcite is present it is converted in lai^ part to sulphate. Galena also roasts io a sulphate, and tends to envelop particles of blende, and to prevent their roastii^. Much of the blende from Leadville, Colo., and other Western States, contains silver, and it consequently often pays to treat the retort-reudues aft«r the zinc has been removed.
There is a loss of silver in roasting that may be given at 10 to IS per cent and also a loss of zinc as dust and volatilization at the final high temperature that may be reckoned at 2 per cent or more.
In roasting zinc sulphides, and especially flotation concentrates, better results in retorting are attained by finishing the roast without an excess of air. Thus roasted, the product ranters and beccnnes more porous, and the sulphur left in the ore is less harmful.
KOASrmO FDIUVACES
The roasting of blende has been performed in hand-rabbled reverberatory furnaces as well as in a great variety of mechanical furnaces. These are described in the chapter on Roasting. The latter are gradually supplanting the former because of the saving of labor. It should be noted, however, that the wear is great on mechanical furnaces that have ironwork exposed to the heat because of the high final heat needed in blende-roastir^ and consequently the types of furnace have been preferred where the rabble is exposed but a short time to the action of the fire, and where iron parts are not exposed or can be water-cooled.
Thus, the Brown horseshoe furnace, where the rabble is drawn through a circular hearth, then allowed to cool, or the Wethey furnace, where the rabble is exposed to the fire but half of the time and the moving iron parts are outside the furnace, have been successfully used in blende-roasting. Of the recent types, the Hegeler furnace has proved most succeaaful for the above reasons. It is a multiple-hearth furnace closed by swinging sheet-iron doors at the ends, and stirred by rabbles drawn quickly throu^ the furnace by means of rake rods, so that the parte are outside the furnace
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The Wedcx Roasitno Furnace S21
most of the time and no iron parte, except the end swingiDK doOTs, ate affected by the fire. The hearths being superimposed make a compact furnace, and the radiation b greatly lessened, eo that there is ectmomy of fuel.
THE WEDGE HBCHAmCAL BLENDE-ROASTING FDRHACB
This, one of the most successful of the blende roasters where it is desired to save the sulphur fumes for sulphurous acid, is shown in Fig, 281.
It is commonly used for an oxidizing roast, the gases carrying about 2^ per cent sulphur, but for use in making sulphuric acid the gas should carry 6 to 7 per cent of sulphur. This is thus accomplished : The blende carrying about 25 per cent of sulphur is self-roasting, the sulphur escaping at the desired strength. Below the fifth hearth, however, when no more than 8 per cent sulphur remains, it needs a fire to roast it. The floor of the fifth hearth is thin and between it and the roof of the sixth hearth is a muffle space where the flame from the firebox at the left enters and is down-drafted. That is, the flame passes through drop-holes, then successively through the sixth and seventh hearths to a ade-fiue leading to the chimney, and in its progress roasting the ore from 8 per cent down to 1 or 2 per cent, the degree needed for properly roasted ore. In the thickness of the side walls is s drop-hole or passage downward debouching into No. 6 hearth. Ore pushed into this drop-hole fills it with a talus at the outlet which is swept away by the rabble of No. 6 hearth. Thus the ore passes down by a sealed opening while the fire gases cannot escape upward, To mingle with the strong sulphur fumes produced above,
Tee Hegeler Fdrbacb
Fig. 282 gives a transverse section, a longitudinal elevation and section and a plan of the fiu-nace, of 75 ft. effective length, and built double. Referring to one side there are seven roasting-hearths each marked B, three fiie-hearths A, wid a hearth C, for preheating the air for the roasting hearth. Thus the three lower roasting-hearths B, are heated by the fire-hearths, constituting muffles, bo that the fire-gases do not mingle with the roaster gas. In operation, the furnace being at full heat, the ore, to the amount of one charge is dropped from the hopper D upon the upper or No. 1 hearth. By the action of a rake of the width of the hearth, introduced throuf^ the end door at that end, it is pushed along and spread out upon the just^ranptied hearth. The next time the r^e is passed throi^ the ore is worked to the opporate end of No. 1 hearth and through the opening E, to the second hearth B. Another rake propels it throw^h B, and so progresravely through the remaining hearths until, at the lowest one, it discharges throw^ a side-opening into a waiting ore car. The path of
Boasting Zinc Ores
6E0T10M J U "" 1 W-*"
Fio. 281. — Wedge Rooster for Zinc Oiw.
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The Hegeler Roasting Furnace 523
the fiie-gaaes is in an opposite direction, the producer or natural gas being admitted to the lowest hearth A, at one end, and traveling to the other, being partly burned by the admiaaion ctf air at side-ports. The gas then by a by-pasB at the ude goes to the hearth A, ttiere to be further burned, and the combustion is finished on the third or upper hre-hearth A. At the end of this the gases pass into a fiue of their own, and thence to the chimney. Air for roasting is admitted, partly in the lower hearth B, partly in the two hearths above it, and taking a course opposite to that fd the ore in its descent, leaves the top of the furnace at the opening F, into the flue G. It will be observed that the furnace is built double, so that when
Fia. 283— Hegeler Roasting Furnace.
ore is admitted at D, it is also admitted at X on the opposite end, and the same rake works both hearths in succession. The rakes or other moving parts do not remain constantly in the furnace, but a rod is passed through from one end, hooks on to the take and pulls it back with it.
The furnace is used in this country where it is necessary to convert the sulphur gases into sulphuric acid, but on account of the high labor and maintenance costs, is rarely used where acid is not made. The capacity of the furnace is about 45 tons per day, with a coal consumption of 25 to 35 per cent of coal dependent on conditions. The furnace is also sometimes built with regenerative chambers for the fire gas, whereby the fuel conaumfH tion is reduced where the high price of fuel makes this imperative.
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Roasting Zinc Ores
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The Merton Roasting Furnace
Various Furhaces
Other furnaces used for the roasting of blende where acid is not made from the waste gases, are the Zellweiger and the Ropp, both of them straight line sii^e hearth furnaces of the Brown-O'Hara type. In European practice the Merton and the Ridge furnaces aie somewhat generally adapted for blende roasting, although on account of cheaper labor, hand roasters are somewhat common there.
The Herton Fdrkace
This b a muffled furnace, having six muffles, where the ore is roasted, and between the floor of one and the roof of the next a space transverse by the fire-gasea. Referring to the longitudinal section, Fig. 384, the ore, delivered by a feeder (not shown) at the left end, falls upon the No, 1 hearth, and by means of eight rabbles is handed along to the other end where it drops upon No. 6 hearth and escapes by a drop-hole into a car standing upon a track below. It will be seen that No. 3 hearth is muffled neither above nor below. The fumes from the three top-hearths go to the sulphuric plant, the fire-gases entering the muffled spaces between hearths 4 and 5 and between 5 and 6 are down-draf t«d to the chimney, while the muffle gases between 1 and 2 go direct to the stack. In this way the ore is soon heated to ignition, and later, on 5 and 6 is strongly heated to give a low sulphur product. It will be noticed that the mufilc roofs are so low that the gases traveling over the hearth continually sweep along the escaping SO2 gas. On hearth No. 6 there is no muffle below so that the ore becomes cooler and in so doing heats the entering air.
The Ridge Furnace
Fig. 285 shows three sections — one a central longitudinal, one transverse, through B,B, and one as a plant at B,B, cutting through hearth No. 3. The upper, the drying and preheating hearth, ia open to the air, and the roasting of the ore is done on hearths Nos, I, 2, and 3, passing thence to the cooling hearth. Below hearth No. 3 are flues for fire-^ases and for fresh air, which pass away by a separate flue. In this way hearth No. 3 is maintained at a high temperature, being a muffle hearth. After the furnace has been heated to a high temperature at starting, the blende is self-burning, air for the purpose being admitted by the proper side-ports aa shown on the plan. The gas evolved from the burning, and containing 6} to 84 per cent SO3 (free from fire gases) passes by the roimd gas flues, and through a suction fan, to the sulphuric-acid plant- The direction of flow of this gas is contrary to the movement of the ore. There are four vertical hollow shafts carrying rabble arms furnished with rabbles or blades
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Roasting Zinc Ores
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Sulphuric Acid Making 527
set at on angle. The shafts being in motion, ore from the feed is swept along by the rabbles the length of the drying and preheating hearth. It falls throL^ the open drop-hole at the firebox end of the furnace to roasting hearth No. 1, thence along the hearth to the drop-hole delivering to hearth No. 2 and so on to hearth No. 3 and to the cooling hearth. The discharge is at the side of the hearth and outside the bnck wall which protects the gearing from dust and heat.
SUU>HDIUC Acm
This is & by-product of zinc roasting. The sulphur dioxide fumes arising from blende roasting are incieasin^y used for the manufacture of sulphuric acid. This is done not only because of the profit arising from this manufacture, but because escape of these fumes into the atmosphere results in damage to health and to vegetation, so that such disposal of fumes has been restricted by legislation. Most blende is free from arsenic, and the fume arifong from its roasting makes a superior acid as compared ' with that made from pyrites. Due to this freedom from arsenic the contact process for making acid is much in xise in the United States, though the older chamber process is employed, especially for the Western blendes canyii^ iron and lead. One may note in connection that it has been found to be an advantage to roast the blende in one establishment near the line mines and the sulphuric acid market, and to smelt the roasted product at another works where fuel, clay, and skilled labor are best found. To roast blende to the beat advantage the operation should be carried on in a multiple-hearth muffle-furnace where one can be sure of a regular supply of sulphur dioxide, free from the combustion gases, and of suitable grade. The objection to band-roasting is production of an irregular gas due to uneven firing and stirring, whereas in the mechanical furnace such conditions do not exist. Again, in a muffle furnace the fuel gases are kept separate from the sulphur fume arising from the roasting ore, and these may be maintained at the grade of from 6i to 84 per cent sulphur as best suited for acid making. In 1920 the average price of 60° Baum4 acid was 114 to $18 pertcoi.
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Chapter Xlii
SMELTING ZmC ORES
The 8Meltiho Or Distillation Of Roasted Zihc Ores
The recovery of zinc from the ore consists in distillation of the roasted ore in refractory clay retorts after intimately mixing it with 40 to 60 per cent of its weight of fine coal. The whole is brought to a white heat, which is maintained during an entire day.
Reactions that Occur in Retorting Roasted Zinc Ore.
(6) 2ZnO + C = 2ZnC03. 172,000 97,000 =+57,000
Before ore and coal are charged into the hot retort the mixture is moistened for convenience in charging, the water being promptly driven off by the heat. The light hydrocarbons of the coal come away next; then the iron oxide is reduced to protoxide and part of it to a porous iron or iron sponge. The final reaction (6) is the reduction of the zinc oxide of the ore by the carbon to metallic zinc. The reaction commences at 1060° C, but practically a temperature of 1300° C. is reached. I
It is to be noted that the reduction point of zinc oxide is conaderably above the boiling-point of zinc so that metal so reduced in the retort is inunediatly carried off with the other products of reduction into the cooler condenser. Should the temperature of this condenser rise higher than the boiling-point {circa 975° C.) the zinc will of course fail to condense, and should it fall below the melting-point {circa 418° C.) the metal will be condensed as powder, known generally as bhie powder. This later characteristio is takisa advantage of in the manufacture of zinc powder (»- blue 528
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Zinc Smelting Furnace
Fia. 286, — Zinc-Bmelting Furnace.
powder — now so much used in the reduction of gold from cyanide solutions — by using iron condensers maintained at a sufficiently low temperature. The condensers employed for spelter are truncated hollow cones of fireclay, which fit just inside the mouth of the retort. These are generally made
Dg,:,z..byG00glC
about 18-24 in. long and taper to about 4 in. outside diameter at the smaller end, with walla about } in. thick.
In the United States the Belgian style retort ia chiefly used. These retorts are plain cylindrical vessels, closed at one end, and are made 48 to 54 in. long, and 11 in. out^de diameter. The thickness of the side walls is 1 to 1} in. and of the end about 1) in. The older style of zinc furnace designed for the direct use of coal as fuel, and known generally in this country as the Belgian furnace, is shown in
Flo. 288. — Section of Zino-amelting Furnace (gBS-Gred).
cross-section in elevation in Fig. 286. Fig. 287 also shows two photographic views of this same type of furnace. On account of their high labor cost in operating these furnaces are rarely used to-day. Instead the larger and lower furnace shown in cross-section Fig. 288 and in front elevation Fig. 289 is more generally employed both for the use of natural gas and for producer gas as fuel. The furnaces are built with any number of retorts, generally 288 to 336 to a side where four retorts high. It ia ' common where producer gas is used as fuel to build them five and even six rows high, in which case the furnaces are made 400 retorts and 432 retorts
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Zin&Smelting Furnace
to a side. Where producer gaa is employed fuel producer are located on the end of the furnace, and all the gas is allowed to enter at ihaX end. Air is supplied at intervals, by blowers or fans, through the pipes A shown on top of the furnace and distributed through the smaller pipes B to the front at each section. The gases pass out through stacks located at the end opposite to the producers.
Where natural gas is used for fuel, no producer is of course employed.
Fia. 289.— Front View of Smelting Furnace (gas-fired).
and that end of the furnace is closed up. The gas is admitted at intervals along the furnace with the air, and the products of combustion pass out the stack on one end of the furnace.
Both these styles of furnace are extremely wasteful of fuel, as the products of combustion leave the furnace at the full temperature of the last retorts. Sometimes waste heat steam boilers are loi^ated behind the furnaces in order to recover this extra heat in a useful form, but as the amount of power required around a zinc plant is comparatively small, only
a portion of the waste is so recovered. For that reason and because of the increaaii^ coat of coal, regenerative furnaces are coming into increasing use. In these furnaces the waste heat of the cutting gases is employed to heat brick checkers, which in turn give up this stored heat to the incoming air and gas. In that way a saving of fully 60 per cent of the coal consumed is made, although at the cost of increased labor to some extent.
Fia. 290. — Smelting Furnace in Uperatioa.
One of these regenerative or Seimens' type furnaces is shown in croaasection in Fig. 291, and Fig. 292 also gives in outline the general method of admission of gas and air through the regenerative chambetB into the laboratory of the furnace, and their movement through the other chambers and flues to the stack. Periodical reversals of the gas and air take place at intervals of half to one hour by means of the valves showi., whereby the chcckcrwork-filled chambers are alternately heated, and again give up their heat to the incoming gases.
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Zin08Melting Plant
There are various des^nsof these furnaces, but nearly all work on the reversing principle as above outlined.
Fig. 293 shows in detail a zinc-retort in place in the furnace. It is
SECTION *.* emelting Plant.
Fici, 292.— Plan of Zinc-smelting Pl&nt.
made of fireclay, 4 ft. long by 8.5 in. diameter and with walls 1.25 in. thick. In the figure a is the retort, which rests on a ledge on the rear wall of the furnace and extends just through the thin (4J-in,) front wall. The wall is held by buck-staves c which carry the tiles upon which the retorts rest.
Smelting Zinc Ores
Fia. 2
The whole is firmly bound together with tie-rods. When the ictort has
been charged, the clay condenser b is set in place, in which the sine vapor iseuing from the retort is to condense. As seen in the front view, the space between two buck-fitavee is divided by shelves which form " pigeon-holes," each of which contains two retorts. The retorts having been set in place, the opening around them is bricked up with pieces of brick and with clay. When a retort becomes cracked or otherwise useless, it can be readily removed by breaking away the teanporary
wall, and another retort can be set in the place without distuibing
the adjacent retorts.
Operating The Fohkacb
The roasted blende, or oxidized ore, or a mixture of the two, is thoroughly mixed with fine coal, and moistened with water so that wbm thrown into the retorts it will pack closely. The coal used for reduction is generally a low volatile, low sulphur coal, preferably anthracite. In the western field various coals are used for this purpose. SometiniEs anthracite is used alone, sometimes what is known as " dead coal," a non-coking weathered coal found near the surface in Kansas; and someUmes
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r
crushed coke or coke braize is employed with either, or even a mixture ol all three. The amount employed is generally between 40 «id 60 per cent, dependent on the character and grade of the ore. This fuel is always used crushed not coarser than 1 in., but generally as fine as ^ in. at least.
The amount of ore chained per retort is usually about 60 lb. with a proper quantity of fuel. The amount of charge for a furnace is placed in one or more cars on the tracks shown in front of the furnace, Fig. 288 and £%. 289, and is shoveled directly by means of the scoop, fig. 294, into the retort. After filling the retort an iron rod ^-in. thick is run along the top of the charge next the retort to provide for a vent for the moisture and gases of the charge.
As the retorts are filled the condensers are set in place resting on supports on the plates in front of the furnace. The condenser is then luted
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Operating The Zinc Furnace 535
or loaded around the joint with the retort by a finely ground and dampened mixture of coal and field loam; hence the term " loaming." The open end is loosely filled with a handful of a mixture of coal and cbai^ or waste material, in such a way as to prevent the flowing out of the zinc, yet permit the escape of the reduction gases.
After the ore is charged the heat of the furnace is gradually raised so as to drive off, first the water added tfl moisten the charge, then the volatile matter of the coal, and the carbonic acid, if any in the ore. Finally after about two or three houra it is raised to the heat of reduction of the zinc.
The penetration of the heat from the outside of the retort, to the inside of the chai^ being progressive, of course these periods of the process necessarily overlap each other, so that before tiie center of the char^ is fully dried out the gases are coming off that part of the charge in' contact with the walls; and before these latter are completely expelled metallic sine is being given off. This causes a dilution of the zinc vapor in the earlier stages so that then the condensation of the metal is incomplete, and much blue powder is formed as well as zinc vapor lost. From this period the temperature of the furnaces outside the retorts is maintained at not less than 1200° C, increasing to 1400° C. towards the finish of the operation.
After the charge has been in the furnace for ten or twelve hours there is generally sufficient metal in the furnace for first metal draining. The metal drawer brings underneath the outlet of the condenser a cast4ron ladle swingmg on a crane, and by means of a special tool breaks out the stuffing of material in the mouth of the condenser. Fart of the metal runs out, and the balance, with more or less oxide, blue powder and portions of the charge which are carried out, is scraped into the ladle by the same tool. This tool is a cast-iron button about 2 in. diameter riveted on the end of A-in. iron handle. The metal is poured from the ladle into cast-iron molds. The resulting slabs are about 12 by 18 by 1\ in. and weigh about 60 lb. each. After the metal is drawn the condensers are closely stuffed as before, and the operation continues without interruption. At the end of about eight hours more the furnace is drawn, and agfiin at the end of about four hours, when the operation is finished.
The condensers are then chiseled loose from the retorts and moved to one side. The loose unworked charge and oxide around the mouth of the retorts are scraped out, and the furnace plates and floor are thoroughly cleaned up. The cleanings, t<^ther with the oxides, skimmings, etc., made during the process are put aside to be charged again, and the furnace ia ready for cleaning out. '
The covers of the openings in front of the furnace are now removed, and the operatives scrape out the residues. These residues flow throu^ the openings into the cellar below, where generally cars are ready to receive them. Sometimes other methods are used to remove the bulk of ths res-
Cooglc
dues iTOm the retorts, but scrapers, or bumpei^ bo called, have to be employed for a final cleanup. The castings and floor in front of the fui^ nace are now swept clean of the spent residues, the covers of the cellar holes are replaced and the furnace is ready to recharge. The operation from this point is as described above.
After the furnace is cleaned out any broken or corroded retorts are readily discovered and are removed. This is done, and they are replaced by new retorts which have been brought up to red heat in a kiln for that purpose, and without being cooled down are pushed into place in the hot furnace. When the retorts are in place they are closed into the fnmaee by a fireclay partition which fits closely around the mouth and closes the opening to the furnace so as to retain the fire.
The heat of the furnace is allowed to fall slightly during the time of charging and changing broken retorts, but at no time during the campaign, which may last five to seven years, is the furnace allowed to cool off.
It«torts to withstand the high temperatore and corrosive action of the charge are made of the most compact and durable material. Tbt material consists of a mixture of " chamotte," " grog," or " cement," of burned fireclay, firebrick, or tile free from slag. It is ground to about 6-mesh size and mixed with an approximately equal amount of raw fireclay. The mixing is done in a pug-mill, using water to form a stiff mud, which is allowed to stand some time covered with wet sacking to season and to develop the plasticity. It is again put through the pug-mill, and 6nally made into retorts in a hydrauUc retort-making machine under the pressure of about 3000 lb. per square inch. A machine of this kind makes twenty retorts or more per hour.
The success of the retorting operation depends upon the durabiUty of the retorts, and for this reason a careful selection of the clay is the first necessity. With one or two exceptions, all the zinc smelters in the United States use a fireclay found in large quantities in the Mississippi Valley, chiefly at or near St. Louis. It is probable that the good results obtained with retorts made from this clay are partly the result of a familiarity and knowledge of its qualities for the purpose, as the attempts made to >we other clays of apparently superior physical and chemical properties, have for the most part resulted in failure. The analysis of this clay, generally known as " Cheltenham " clay, is about as follows, on an air-dried sample:
and loss on ignition, including water and organic matter about 15 per cent
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Retort Making 537
Condensers. — CoDdensera are made of less refractory clay than letorte. They are not subjected to high temperature, but must withstand much handling and severe treatment. They last eight to twelve days, and coet 3 to 4 cents each.
Drying the Retorts. — The finished retorts as they are removed from the machine are placed in vaults or compartments holding 500-1000 retorts, according to size of works. These vaults are provided with steam coils beneath the floor, for heating and drying. In general the temperature fA the vault is kept low during the first week or more, to allow slow evaporation of the added water, then is gradually raised during successive periods until the retort is thoroughly freed from moisture. The period of seasoning may hardly be less than four weeks, and may better be prolonged for at least twelve weeks. It is considered better that it should be a slow operation, but requirements in this respect differ with different clays.
Annealing. — ^Af ter the retorts are thoroughly seasoned and dried they are available for use in the furnaces as required. Each day the number of retorts experience has shown to be required are placed in a small furnace or " temper-kiln " cold, and the heat gradually raised so that first the combined water b removed and later the heat of the kiln is brought up to fuU redness. About twenty-four hours is usually required for this purpose, and at the proper time, as required in the operation of the furnace, these retorts are taken while still red hot and placed in position in the smelting furnace.
LOSS m THE PROCESS
The losses in smelting of zinc occur, as may be expected, in every process to which the ore is subjected.
In blende-roasting there is a mechanical loss from spilling of ore, and from flue dust, and a metallurgical loss of zinc in the form of fume due to the volatiUzation of zinc.
In smelting the roasted blende or carbonate silicate or oxide, there are again mechanical losses in spilling and dust, but the more serious losses are those of the metallurgical process itself. These may be summed up as: (1) Infiltration of zinc through the retort and absorption of zinc by the retort itself; (2) loss in fume from the condenser due to tmcondensed zinc ; (3) fume or zinc vapor remaining in the retort at the conclusion of the process; (4) zinc remaining in the residuum after removal from the retort.
In the roasting of blende the losses will amount to at least 1 per cent and will average 2. In the case of fine ore, dust tosses may increase this proportionally.
In retorting roasted blende which averages 40 per cent zinc before roasting] the losses will aven^ 16 per cent; where the zinc tenor is as
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high as 50 per cent tlie loss will approximate 13 per cent and when 60 per cent material is handled the average loss will approxiniate 10 to II per cent. Based on the original imroasted material, the total of all losses will be approximately for 40 per cent material 18 per cent, 50 per cent will be 15 per cent, and 60 per cent about 12 to 13 per cent. Special conditions and special characteristics of ores will of courae modify these figures.
In smelting carbonate and silicate ores, which average lower in zmc content, the per cent of loss will be less. These ores for the most part have ranc contents of 30 to 40 per cent and the zinc loss will be from 12 per cent for the better grades to 17) or even 20 per cent for the poorer.
Cost Of Smelting
The cost of the smelting of blende of course varies much with different locaUties. The large natural gas fields found in the Kansas-Oklahoma territory enabled cheap operations to be carried on at points so situated with respect to the most important zinc deposits of the country that the freight charges on the ore to the works and on the metal to the point of consumption, were reduced to a minimum. The costs in that field, therefore, have been and are to-day exceedingly favorable, considering the character of the operation. Exclusive of gas cost these chaises in 1912 are about as-follows:
Per Ton Bleoik.
Unloading, crushing, drying, and sampling SO. 35
Roasting, labor and repairs 0.50
Smeltini. Per Ton Ronalad.
I«bor, direct $4,00
Reduction fuel 1.80 ,
Retorts, condensers, etc 0.60
Repairs, maintenance, power, etc 0.73
Charging, etc 0.35
General labor 0.20
Superintedance, office expenses, etc 0.75
S8.45 Roasting loss 13 percent - 1.10 7.35
Total, exclusive of gas $8.20
The cost of gas varies with the conditions surroxmding the plant. In the early days of any gas field the cost is apparently very low, but increases rapidly with the exhaustion of neighboring fields. Probably a fair average cost of gas per ton of blende for both roasting and smelting would be $2 to $2.25 per ton. This added to the other costs makes the total cost average from $10.25 to $10.50 per ton for the whole operation.
In the coal fields the cost of fuel b h^her, but owing to the locatioD
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Prices Of Zinc Ores 639
generally chosen for these plants u^i^ coal for fuel, there is a good market for sulphuric acid at such points, and the manufacture of acid as a byproduct reduces this cost to a point generally much below the cost at natural gas-using plants.
PRICE OP ZmC ORES AND SMELTERS m 1919
Mississippi Valley ores are bought as at Joplin, Mo., at a quoted price, on a basis of 60 per cent zinc contents. As this varies up or down, SI per unit is added or subtracted from the base price. Calomine is bought on a 40 per cent basis. We may quote for a certain blende $47, and for caltunine 930 per short ton.
Western Zinc Ores. — On a 40 per cent bams we may have, for example, $20 per ton paid for a calamine ore or S13.50 for a sulphide. A variation of II per ton per unit is made up or down. Besides this 65 per cent of the sine contents is added or subtracted as the market price of zinc ris^ or falls.
Iron over 2 per cent is penalized at 11.50 a imit; arsenic or fluorine are not permitted. The gold and silver may be paid for at 65 per cent of their market v^e. The above figures are based on zinc at 8 cents per pound.
Zinc — Quotations m the United States in 1910 are given in cents per pound, thus: 4.60 cents, St. Louis; 4.75 to 4.80 cents. New York. The London market is quoted at 9.15 shillings for good ordinaries (ordinary brands) and 20 shillings for specials (the purer zinc). St. Louis is near the zinc-producing district of Kansas, Missouri, and Illinois, and hence has a lower price for spelter than New York.
The Boropean Price of Zinc Ores. — The vahie of a zinc ore depends upon its content in zinc and the absence of objectionable impurities, such as iron, manganese and lime, which form Fusible slags, and mcrease the corrosion of the retorts, or of lead, cadmium, arsenic and antimony, which contaminate the spelter and ao lower its market value.
For many years lai^ quantities of zinc ores have been sent to Antwerp, Belgitmi, and to Swansea, Wales, from Sardinia, Algeria, and Spain, and of late from Colorado, via the Gulf ports of Galveston and New Orieans.
Based on oRS of 46 per cent zinc and upward the price of the am, with costs, insurance and freight p^d, was the London price less 8 units and 95 per cent of the assay value, and less a returning or treatment-charge oi perhaps £33 per ton.
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Chapter Xliii
Zinc Refinino
Refining is for ibs purpose of converting low-^rade spelter into grades suitable for high-grade brass. For such grades from 5 to 15 cents per pound over prime Western spelter was pwd during the Great War.
Grades Of Zihc
Electrolytic spelter of 99.9 to 99.95 per cent is the purest made. Aside from this the great bulk of spelter made by distillation, and called virgin spelter, has been divided into four grades, as follows:
A. High
B. Intermediate
C. Brass special
D. Prime Weatera
In distillation lead is volatilized and carried over into the spelter. For the grade " brass special " the lead in the ore should be below 1 per cent. Iron makes less trouble, and may occur in ores up to 10 to 12 per cent. Cadmium is even more easily distilled than zinc and so first^draw zinc contains the most of it. It is not considered to be detrimental in small proportions.
RediatiUing. — To prepare for this the ordinary ore-smelting furnace has its lower row of retorts removed and the butts of the upper rows are placed upon the shelves of the next lower row. This gives the retort an inclination of 8 to 10 in. in its length and permits it to hold a bath of molten spelter. Since in redistilling a lower temperature (950" C.) can be carried than in smelting, the flue checkers are opened and the combustion gas is burned imder natural draft. There results a thin flame, a more xmiform heat and better operating conditions. The condenser used may be an ordinary one, having a dam block in half the larger end; or ^ain a handmade cone condenser may be preferred. This smaller end, 7 in. diameter with a tile dam, is luted into the retort, while the lai^er end, 10 in. diameter
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Zinc Refining 541
is cloeed with a fireclay plate with tap-hole openings, top and bottom. These openings are stuffed up during distillation. The low-grade spelter which is to be refined is cast into sticks 20 in. long by 1) in. square. Four or five of t^ese make a charge for one retort, these being charged immediately after each drawing. The sticks are inserted by the top opening. This ia etuffed, and in five or ten minutes the sticks are melted and in process of distillation. Drawing ia done every six hours. To do this the top opening is spiessed or pricked open, in order to relieve the internal gas pressure, the scratcher is inserted in the bottom hole and the contained metal and blue powder drawn into the ladle. The redistilled spelter is cast into plates and taken to a reverberatory equalising furnace for recast-ii^ into plates of uniform high-grade spelter.
As the bath of metal in the retort becomes enriched in lead and iron it must be removed. This is done by omitting the charge for say, twenty-four hours on one section of nine retorts, and the next day taking down the condenser of this section and scraping out the metal called " bottoms," using a lai^ scraper. In esse of a leaky retort all is removed and the retort at once replaced. The leady bottoms drawn into a ladle are cast into plates and taken to a remelting furnace where any excess zinc is separated. The lead, tapped from this furnace and carrying 1 to 2 per cent zinc, is st^d to lead refinenee. The blue powder, skimming etc., from the redistilling, remelting and equalizing furnaces are returned to the ore furnaces for resmelting. All scrap metal is sent to the remelting furnace to be recast into sticks of recharging. Five men operate the two furnaces of the block.
To refine spelter, a furnace resembling that shown in Fig. 241 is used, but the reverberatory firebox is in two parts, and provision is made to charge the spelter close to the bridge. It holds 30 tons of spelter when full, and in it 10 tons can be refined in twenty-four hours. The metals separate into layers. At the bottom is the lead ; the iron forms with the zinc and part of the lead, a difficultly fusible alloy that floats on the lead, and uppermost is the stratum of pure zinc. By means of an iron rod inserted into the bath, layers are distinguished, the zinc being soft, the iron-lead-zinc alloy (called " hard zinc ") being mushy, and the molten lead at the bottom soft. The underlying lead is removed weekly. A cylinder or pipe closed at the lower end is sunk below the lead layer. The plug is then knocked out and the lead, rising in the cyhnder, is ladled into molds. The zinc of the top layer is ladled out daily into molds, and it retuns 1 to 1.25 per cent lead. The hard zinc layer is removed when opportunity offers. To do this the zinc is ladled out first, the lead is next removed, and finally the mushy mass of ferruginous metal is removed with ladles perforated so that the lead drains off. This hard zinc is sold for the manufacture of Delta or Sterro-metal.
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Tbb Db Sauixes Redistillation Method
A furnace is used like one side of an ordinary furnace block. The letorte are inclined 7 in. in their length and extend 4 to 5 in. through the back wall At the top of the protruding back is an opening for ohaqpng the retort with molten spelter. At the bottom is a small tap4iole for removal of the leady bottoms. Both openings are ti^tly closed with clay except when charging or tapping. There is an ordinary condenser properly clayed to the retort. Each retortKtistilling furnace of, say, 200 retorts, is served by a 25-ton remelting furnace and a 25-ton equalising furnace.
Low-gnkde spelter for redistillation coni^sts <^ second or third-draw metal, and contains 1.5 to 3.0 per cent lead, 0.03 to 1.0 per cent iron and 0.03 to 0.07 cadmium. The redistilled spelter will average 0.10 per cent lead, 0.01 per cent iron and 0.04 per cent cadmium, a better grade than " intermediate," as given in the table.
Befihihg Sevlter Withodt Redistillatioit
The principle of this method consists in remelting low-grade spelter in a TGverberatory furnace with a reducing flame, and letting the molten bath stand tmtil the metal separates into layers according to the specific gravity of the different metals, the lower part of the bath consisting ol a leady zinc and the upper part of spelter nearly free from lead. The lower layer is then tapped, or removed otherwise. The separation or refining must be done at a temperature near the melting-point of zinc, since the higher the temperature the more persistently does the zinc retain lead. Under the most favorable circumstances the lead content of the spelter is reduced to 1 to 1.26 per cent.
Due to the demand for a pure metal for the making of cartridge brass during the great war, electrolytic zinc was held at a premium of 2 or 3 cents per potmd over the then high price of other grades. This gave a great impetus to developing a successful electrolytic method for its manufacture.
The process consists briefly in roasting the ore, dissolving in dilute sulphuric acid, filtering the solution, precipitating the other bases by means of idnc-dust, filtering to obtain a solution containing idnc only, and precipitating this in metallic form by electrolysis.
Roasting of Zinc Sulphide to Oxide and Sulphate. — Zinc sulphate can be formed throw^h any one of the following reactions:
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Electrolytic Zinc 543
(12) ZnO+SOa+Fe203 = ZnS04+2FeO (fr).
There ia connderable evidence that the first reaction is responnble for most of the sulphate formed. The only gaseous reagent is oxygen and there are no gaseous reaction products, therefore, the oxygen concentratitm alone should mainly determine tlie amount of sulphate formed. Keactions (11) and (12) involve two gaseous relents, so that the amount of BUlphate formed will be determined mainly by the product of the concentrations of oxygen and sulphur dioxide.
Ttie ore, of SO^nesb size, is a blende concentrate of 25 per cent sulphur. It is roasted in a Wedge roaster, being given a close oxidizii^ roast that brings it down to 2 per cent sulphiu*, yielding a product of zinc oxide and sulphate. This, after cooling, is stored in feed-bins, whence it is drawn off into agitating vats. Here it is treated for twenty-^otu-hours with dilute sulphuric acid. The acid dissolves the zinc oxide and small amounts of copper and cadmium also present. The pulp is now passed over a classifieT which removes the sand, the slime then going to an OUver filter. Both the sand and the slime are stored and are sold at a profit to the smelter as containing silver and gold and a little lead.
There rcnuuns the clear solution, which is passed to another agitator. Here it is treated to a small additirai of granulated zinc made in the melting house from zinc produced by electrolysis. After a prolonged treatment, in which the zinc dust precipitates all the other bases, the product is pumped through a closed Sweetland filter, where the zinc-dust and the bases aro removed, leaving a clear solution containing zinc tmly. This solution is stored in vats and is drawn off as needed to the electrolytic tanks of the tank-house, arranged and operated precisely like the tank-house of copper refining. Each tank contuns eighteen anodes and nineteen cathodes, the current being in parallel with a tank-resistance of 0.4 volt. The anodes are of lead, 21 by 36 in., while the cathodes are of aluminum 24 by 36 in.
When the zinc has been deposited to the depth of i in. on the cathodes these are removed from the tank, washed, and the zinc is stripped. This is melted in a reverberatory furnace, and melted into commercial plates, 10X16X2 in, A small portion of the molten metal is granulated, it being placed in a small reservoir whence it runs in a stream the size of a knitting needle so as to be caught by a horiisontal Eur jet. This blows it to powder and it is caught within a sheet steel bin or chamber.
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PART vni
Plant, Equipment And Their Costs
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Chapter Xuv Location, Bquiphent And Brbction
Location Of Works
A mine has practical value only when permanence ia reasonably assured. Then, asmgle unit of a mill may be erected where the process that has been chosen may be perfected ; after that other units can be added.
Mines Works. — If a mining company builds a mill for the treatment of its own ore, it is usual, to save freight or hauling expense, to place t&e plant as close to the mine as the securing of a suitable site and water-supply pennies. In case of a smelting works where flux and fuel is to be brought in and a heavy product shipped, then, in addition to a good site and water supply, neamess to a railroad is to be considered.
Custom Works. — This can judiciously buy ores from neighboring mines, but needs a site convenient to the chief source of supply and to the coke and fuel that it must use. For such a plant a point should be chosen where several railroads give rise to competition in freight rates and where labor is abundant. Low freight rates, abundant labor, and low money rates combine in locating a custom works.
Iron and Steel PlantB-^Thus iron ^nd steel manufacture has centered about Pittsburg, Pa., because coke, coal, and natural gas are abundant, and because a good market is found tiiere for the products. On the other hand, the iron smelter at Pittsburg miiat pay for freight from mine to furnace, $2.25 per ton, and must carry a large supply of ore to last throu(^ the winter months when navigation is closed. The United States Steel Corporation, the lai^est manufacturer of iron and steel in the world, has erected a plant near the iron ranges at Duluth, Minn., for reasons shown below.
Effect of Water Caniage. — Vessels carrying iron ore to Lake Erie ports can retxmi with cargoes of coke or coal to supply the Duluth furnaces, which then have a local market for their pig-iron, and do not need to " stock up " with a winter's supply of fuel. Figuring roughly that 2J tons coal, made into coke and into producer-gas, is required to make a ton of steel, there is a slight advantage, as to fuel, in making coke in by-product ovens at Duluth, Minn., and using gas-engines which utilize the blast>> furnace gases to the best advantage. Nearly two tons of iron ore musttw Bent to Eastern furnaces to produce this one ton of steeL
648 Location, Equipment And Erection
Zinc WoAs. — It is aeen from the coet of producing zinc, that 3.5 tons of coal are needed per ton of ore. Thus it is cheaper to convey ore to fuel, thaa coal to the mine where ore is produced. Near Jophn, Mo., there is ore and also fuel; we expect, therefore, to find the anc-smelting woriis working there to the best advant^e. The region is made more favorable hy the fact that natural gas is to be had there.
Silver-lead Worics. — ^WitJi respect to silver-lead works using lead as a collector of other metals, the favored places have been found to be railroad centers, such as Denver, Pueblo, and Salt Lake. From 12 to 15 per cent coke is used in the charge in smelting, so that nearness to coalfields is not the all-important condition. On the other hand, ores are available there in proportion favorable to combining profitably with one another. The lead of one ore and the iron of another, being combined, serve the requirements of smelting.
The silver-lead and copper custom smelters carry a supply to last from two to four weeks, but at a mines works provision is needed but for chk to two days' running.
Mills. — In treating ore by milling and cyaniding, the uuount of fuel and other supplies required is small, and hence the natural place for the work is near the mine that produces the ore, provided the extraction, or recovery of the precious metals, is high. When, however, the ore is refractory and the recovery is low, it pays to ship the ore to smelting works tliat guarantee a high extraction.
HATDBE OF THE STTB TO BE CHOSEN Both side-hill and flat sites are chosen. For the side-hill or terraced site, much used for mills, the ore is arranged to advance or flow by gravity from one operation to the next, using elevators less than on a level site. Since ore and mill or smelter products are so readily moved by cars and industrial locomotives which easily reach the higher levels of the works, certain objections, when material was man-handled, can be said to faU away. On the other hand the flat site has these advantages: (1) The first cost of the works is less, ^ce heavy grading and retaining walls are not needed. (2) In the side-hill works the different parts of the plant must be placed in a defuiitc and constrained order to obtain the needed fall, whereas on the flat site one can expand in any direction, that is, the parts that have to be far apart on the inclined site, can be placed close together on a flat one. A building on a flat site can be better ventilated than one crowded against the slope. On a flat site elevators are more nimierous, but they are also convenient for delivering just where you want and at a low cost per ton.
Iron and steel plants, the lai^^t in the world, are constructed on level ground. The ore is tmloaded direct from vessels to the stock pile, using
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Plant Sites 549
grab-buckets holding 5 to 10 tons each. If tmnsported in care, the cars aire loaded in a sunilar manner. The custom is to use hopper-bottom cars, from which the ore drops into the charging bins, and thence by charge-cars is conveyed to the furnace-skip. By the skip it is hoisted 100 ft. to the fumace-top. Many recent silver-lead smelting plants occupy level sites, but the dumping ground for slag is at a lower level.
For iron works Uttle attention is paid to the location of the slag-dump. There is no hesitation in sending the slag, if necessary, a mile away by locomotive to be dumped.
On the other hand at copper and lead smelting works the designer likes to have at least two levels. At the largest copper reduction plant in the world, at Anaconda, the . side-hill site has been chosen. Metallurgical mills are very commonly on steep side-hille.
Hill-sites. — On the imclaimed mineral lands of the Western United States, title is secured from the general Government for a mill-site for reduction works, five acres in extent, either in connection with a mining claim (on a theory that each lode claim is entitled to a mill-site) or as a site for an independent or custom reduction plant. A reductitm company, operating a mill, must dispose of the tailing it produces, and of the water discbarged, not encroaching upon the property of other people, and it is responsible for all damages. A company must not let tailing that, at a reasonable cost, can be impounded, flow into a stream, nor run into waters where liable to interfere with navigation. The right or custom of dimiping on the valueless land of lower mining claims is general, except that the practice must do no damage to the property of owners below.
A reduction company can take up lands for a ditch or flume from \mappropriated public land, and the claim cannot be interfered with by later locators; but the owner of such a ditch or flume is responsible for damage arising from breaks or overflowF.
The same rule holds with respect to roads and traUs. In Colorado, mining claims are subject to the rights-of-way of parties hauling ore over them, but in other States the location gives exchisive control, except that a water, electric, or railroad company can take it under the law of eminent domain by fpving a fair compensation for it.
Damage from Smoke. — The smoke from the smelting works, especially those treating sulphide ore in quantity, dehvers into the atmosphere many tons of sulphur-fume daily, as well as fine flue-dust carried out of the stack by the draft. This diffuses through the atmosphere and is carried by the wind to trees and the crops of the land. If not diluted, it blights vegetation, and naturally the farmers org^ze to secure damages, or to close the works. The question of what to do to avoid the difficulties is a serious one, and to-day when pyrite smelting and extensive roasting of sulphide ores is carried on, the trouble can not be altogether overcome. Thus far
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550 lcx;ation, equipment and erection
the sduUon has conasted in locating the works in places where there is little vegetation to be damaged, or in dischaii^g the fume into the atmoe- ' [4iere from hi^ stacks. It may be said that the latter expedient lessens but does not altogether obviate the difficulties. The metallxu^t must thei^ore give serious conaderation to the matter, otherwise, after erecting and starting the operation of a plant, he may find that he is c<anpelled to close it, to the ruin of the entire enterprise. At tiie Washoe Works, Anaconda, Mont., the smoke from the furnaces is treated xmder the Cottrell system for the removal of all dust and fume, so removing this cause of complaint.
Final Consideration. — Preliminary to building a plant and operatii^ a works, an investigation is made of the process, the requirements of the plant, and all limiting conditions. It includes, besides the general matters outlined above, the questions of supplies, marketa, railroad facilities, freight rates, sufficient and suitable labor not liable to strikes, and reliable civil conditions \maffect«d by revohitions or oppresmon by the government under which the plant must operate.
Next comes the organisation of the operating company and finannng of the enterprise, or obtaining capital to build and operate the plant until it pays the operating costs.
Often the promoters, besides owning the mine for which the reduction works are built, have acquired the necessary real estate and the rights that go with it. ProvisionB should be made for access by railroads, for the necessary trackage, and for the common roads to the plant. Not only must water and power be provided, but rightK>f-way for securing them. If fluxes are needed, then the proper quarries or deposits must be found.
consTRUCTion of plaitt
ConBtmctioiL — Before beginning construction, plans should be fuUy worked out by competent engineers. Coat estimates are made in detul, good materials are accumulated, and the labor-force is properly oiganited. In the design of the plant, provision for duplicate parts is made, so that in case of breakdown no mterruption of operation occurs.
On beginning construction, the hydraulic works, where needed, are put under skilled supervision. This includes the building of dams, reservoirs, the water-power plant, and the transmission line. For a long-distance power-transmis^on line there may have to be sub-stations and a distributing system.
Money must be provided for the salaries of officers of the company that are to receive pay during the period of construction, and all money expended must bo accounted for, and cost-records kept by a skilled accountant. The money needed for legal expenses, general expense, traveling expense, and all expenses incurred during construction must be included.
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Chapter Xlv
Accessory Equipment Of Plants
Equqnneot. — This includes the machinery, funiace-toolB, and appliances used in operating, but excludes land, buildings, and trackage. Labor-saving machinery, when reliable, effects a saving in costs, but it is remembered that this saving must not sacrifice the efficiency of operation. The question " how much " often arises, and we may even come to the conclusion that it is not desirable (considerii^ the cost of installation) to put in the labor-saving appliance.
Intermittent Handldio Of Materials
For handling on one level, 100 tons or less of material daily, especially where the ore is to be distributed to various places, one or two-wheeled buggies, or barrows, Fig. 295, on a good floor, have been found to be economical, elastic, and low in first cost. For small quantities the metal-lui^t is not led into installing machinery, for he finds in practice that it effects no
saving. For lai^ quantities barrows or bug- p,„ 295.— Charging Buggy., ^es may be used, or hand-propelled tram-cars, as in mining. For still larger quantities, power propelled cais are used, that can be handled also on up-grades and sent from level to level.
The advantage of this method of handling is that loads can be sent over trestles, above bins, can be raised on elevators, and where they have suitable wheels, can be run over Sootb or even upon the ground.
Industrial Locomotives
These may be operated by steam, by electricity or by compressed air.
The Steam Locomotive. — Fig, 296 is a view of one of these suited to the handling of industrial cars.
The Electric LocomotiTe. — ^We show in Fig. 210 an electric trolley system, suited to the handling of slag and matte on the dump, and indeed for cars also. Fig. 298 is an electrically operated charge car, and one is
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Accessory Equipment Of Plants
Fio, 296,— Steam Locomotive.
FlQ. 2B7.— Gasoline Locomotive.
FiQ. 298.~E)ectric Storage Batteiy Looomotive.
Fia. 299. — Compreeaed Air Locomotive.
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Industrial Cars 553
used at tbe iron blast-fUmace, F^. 154, where it is in fact a movie^ weigh-scales, so that the items of the charge are weighed by it.
The Gasoline Locomotive. — These are made of sizes up to 20 tons and of any desired gauge. Fig. 297 is an example of one of them.
The electric stor^e battery loc<»uotive, Fig. 298, is a view of such a locomotive. It should be observed that with it there is no need of a trolley line, and that the machine will traverse the entire yard trackage.
Compressed-air Locomotives. — Air ctonpressed at 800 lb. per square inch is drawn off from an air-pipe line at a convenient point through a valve and hose into the air tank of the locomotive. The supply will run the machine for several short trips before it must be replenished. At the Washoe plant of the Anaconda Mining Company thousands of tons are thus handled daily, and indeed the plant is an admirable example of how a ^de-hill site becomes effective when locomotives are used. ITg. 299 shows one.
HtDUSTRIAL CARS AND HOISTS For the transport of material about the works a variety of cars are used as shown below.
The Rockn-side I>uiiq> Car. — The Fig. 300 is a view of one of these
Fio. 300.— Side-dump Car.
locked ready for a load, also in dumpb^ position. It delivers outside the track either to a dxmip or into an ore bin.
These may be provided with a brake, may be built with a scale for wei^iing the contents of the car, or may be power-driven, being provided with an electric motor. Some agfun have rubber-tired wheels for operation about the plant without the use of tracks.
Hopper Cars. — These (see Fig. 301) are used for transferring calcine from a roasting furnace to the hoppers of a reverberatory furnace or of coal to the coal hoppers. The bottom opening has a drop door for the dis- (^harge of its conttaits.
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554 Accessory Equipment Of Plants
Transfer Cars. — ^This is useful for shifting a charging machine frtun one Bide to other of a furnace, or of a charging vat (see Fig. 302). The machine is loaded upon it from one Une of track and the transfer car takes it over to the other and parallel track.
Side-discharge Car. — The rocker dump-car discharges to one aide, the side discharge car Fig. 303, at both sides and beyond the track. It is
Fia, 801.— Hopper Car. Fig. 302.— IVanafer Car.
well suited for large loads emptying from an overhead track into tnns below. The one here shown is motor driven.
Hoists. — Of these, the commonest about reduction works is the platform elevator, Fig. 304, which takes buggies, wheelbarrows, or tram-cars from floor to floor. It may have a platform of a size (6 by 6 ft.) to receive two cars or wheelbarrows at a time, and it raises a one-ton load 60 ft. per minuto. They are often nm in balance, but it is better to have two independent counterweighted platforms. Necessarily, time is lost in loading and tmloading, so that the estimate of the capacity is 25 tons hourly.
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Track Crane 666
Skip Car (Fig. 305). — ^This is used for hoisting material by a Bteep incline-track for chargii^ to a furnace. One is shown as part of the equipment of an iron blast-furnace, Figs. 153 and 154. It has a bale for attachment of the hoisting rope.
The capacity of the skip is 2 to 5 tons of ore or half the quantity of coke, and the skips are run in balance. It takes thirty-four seconds actual time for raising, dumpii^, and retiuTi-
ing the sJdp to pit; but the total time includ-pm go.,.— sjtjp Cw.
ing the waits is four minutes, this furnishing
the supply to a furnace producing 350 to 500 tons of pig iron daily from a total burden of 1150 to 1650 tons.
GRABS Ain> EXCAVATORS
Grabs. — In large establishments hoisting rigs are used that are provided with lai^ clam-shell buckets or grabs. They take 6 to 10 tons of ore at a time, and are used for unloading vessels, and for transferring ore
FiQ. 306. — Track-crane and Grab-bucket.
to stock-piles for storage, or to the furnace storage-bins, as desired. It is noticed that the movable frames or bridges are made heavy to cany the lai^ loads safely.
Telpherage refers to the transport by a monorail about a plant, and wherever such a rail can be run there material can be transported. We
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Accessory Equipment Of Plants
give the details of a telpher for the transfer of matte pota from the furnacee to the yard.
The TraveUng Crane. — Fig. 308 gives in elevation a traveling crane as commonly used. It is for hunrlling ladles and converts as described
Fio. 307.— Telpher.
under copper-converting and for Bessemer and open-hearth practice. For handling materials inside a building it is coming into general use. It is operated by electricity, and moves in any direction, horizontally or vertically, over the floor of the building commanded by it.
Fia. 308— Traveling Crane.
and it avoids obstacles on the floor. Provided with large mushroom-shaped electro-m^nets, it is now used to unload pig iron or handle steel sheets weighing a ton or more, and by using magnets no time is lost as in older methods in pasmg chains around objects to be lifted.
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Belt Elevators
coNTnnTous eaudung of materials
Madiines of this kind carry a distributed load, so that the Bub-stnicture upon which they rest is light compared with one upon which the load ia concentrated as in a car. They deliver material contmuously, and no time
Fig. 309.— Bclt-elevatore.
is lost in loading and unloading. Intennittent conveying, on the contrary, if we increase the load of the skip or bucket, becomes slow and awkward, whereas in the continuous conveyor it is possible to increase the capacity by widening the conveyor and providing the corresponding^y ipcreaeed feed.
We divide continuous machines into elevators, conveyors, and conveyoMlevators.
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Accessory Equipment Of Plants
Pia. 310.— methods of Feeding Elevators.
Fia. 311.— Single-strand End-lees-chain Elevator.
Fia. 312.— Double-fitrand End-lesB-ohaiu ElevBtor.
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Belt-Conveyors 559
Elevators are used for vertical or nearly vertical lifting. The belt elevator, Fig. 309, is of this type, and consists of an endless belt having sheetrfiteel buckets, attached by dat-headed elevator-bolte at 18-in. intervals. To allow for the stretching of the belt, the lower pulley shaft is car-
FiQ. 314.— Discharge at Head of Conveyor.
ried in take-up boxes, by which the shaft can be raised or lowered. The lower pulley is enclosed in a boot, the ore delivering into the buckets at the rising side at the left. Ore not caught l^ the buckets falls into the boot and is there scooped out by buckets, and is delivered to the discharge spout by centrifugal action as the buckets pass over the top pulley.
Fig. 311 represents a ^gle-strand endless-chain elevator. The
660 Accessory Equipment Of Plants
chain is carried by head and foot sprocket-wheels with sprockets spaced to take links of the chain. In this case the " take-up " of the aingle-stiaiid elevator is carried at the boot ; for the double-strand one it is at the movable vppet shaft in the left. The ote Bpills into a chute between the two upper
Fig. 315. — Movable Tripper Disch&rgiDg, pulleys, and in this way the elevator can run at the low velocity suited to the type.
The Wonn or Screw-conveyor. — This is convenient for delivering crushed ore or pulverized coal short distances, and it thoroughly mixes the ore conveyed. Fig. 313 represents a screw-conveyor delivering ore from the trough along which it has been conveyed, into another at right an;^. The ore drope from the first U> the second, and b conveyed by the screw
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liELT-CONVEYOHS 561
in the second, shown at the left. The bottom of the trough ia lined with smooth fiheet^teel bent to conform to the wonn or screw. A screw-conveyor is shown m Fig. 268. The disadvantf^es of this type of conveyor are, that much power is needed, and that the ore grinds on the conveyor, resulting in wear.
Belt-conveyois are used for the horizontal transfer of materials, and can be modified easily to carry up an incline. Of all conveyors, the belt-conveyor is most widely used. To give it capacity, it is troughed by running on puUeys that raise the edges of the belt forming a shallow troU{^ (see Fig. 315. The amplest form is an endless belt running over end-puUejrs, the load being fed at one end, delivering into a chute or into a bin at the other. The conveyor carries a load not only on a level, but on as steep as 24" incline. The capacity is lai^ and the conveyors are simple and durable. A 12-ui. belt, traveling at the rate of 150 to 350 ft. per minute, delivers 10 to 35 tons per hour. A 24-in. belt, traveling at the extreme
Fig. 316. — Incline to Level with Movable Tripper.
velocity of 600 ft. per minute, has a capacity of 250 tons per hour (rf crushed ore, and requires 6 H.P. per 100-ft. length, the power needed varying with the length of the belt. When the ore ascends an incline we add the power for lifting the load. Fig. 314 shows a belt in action dehvering its load over the head pulley into an ore bin.
The Movable Tripper. — It is desired at times to dehver the ore into iMns situated at different points along the belt. This is accomplished by uang the movable tripper shown m Fig, 315, which also shows the belt loaded with ore. To discharge the ore the belt goes around the upper pulley, as shown, then around a second one just below, and continues the course to the front end-pulley. The ore shoots from the belt into spouts, that deliver on either side of the track upon which the tripper moves into one of the bins below. The tripper can be moved on its track and set to deliver to any desired bin.
At Fig. 316 we show a tyincal installation, in which a feed hopper at the right delivers to an inclined belt that changes to a level one, the horizontal part having the movable tripper. Beneath the belt ia an idler pulley by whose vertical movement the slack of the belt is taken up.
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562 Accessory Equipment Of Plants
Endless-chain Conveyors. — ^These are much used, since they conv^ ore not only on a level, but vertically if necessary. Being entirely of metal, they successfully convey hot materials.
Kg. 317 represent an endless-chain conveyor, consiBting of a aeries of plates or " flints," attached to a double endless-chain carried at each end by sprocket-wheels like the double endless-chain elevator, Kg. 312. The ore, drawn from any desired storage-bin as shown in the figure, is pushed up an incline by the moving flints in a fixed steel-lined trough, and is taken by a double-strand endless-chfun elevator to a floor above. If desired, slides may be provided in the bottom of the trough. When the slide is opened, the ore drops into the deared bin beneath.
Sometimes, in place of flints, a continuous series of buckets or trays is
used. These overlap so that the ore cannot drop between them. They operate upon the principle of the Heyl and Patdson pig-casting machine. Fig. 162, Indeed, chain conveyors lend themselves to a great variety tA applications, as the examination of a catalogue of elevating and conveying apparatus will show. The chief drawback to them is that they have numerous joints to wear, and that the troughs, flights, or buckets are subjected to serious wear. They must run slower than the belt-conveyor.
In Fig. 74 (elevation) the EMwards roasting-f umace, we have an examine of a swinging push-conveyor. The flights aie bladed and so hinged from the vibrating carrying-beam as to swing over the ore in the conveying trou^ on the backward motion, but to push the ore along when moving forward. As is seen, the bottom of the trough is provided with slides to deliver ihe ore where it is needed.
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Chapter Xlvi
ORE STORAGE AND StPPLY
Provisioh For Supply
Ore may be stored upon the ground or a floor, sometimes fenced in to fonn a ground Ixn. ' Often for-convenience in discharging into a car or c<Hiveyor a bin may be set high enoi^^ to discharge into a car or upon a conveying belt. Bins may be lai^ enough to give a day's supply or less, and are then called " feed bins " ; or there may be a row or series of them for the storage of various materiab needed in the milling or smelting operations, and enough for a number of days. At the luge iron works, receivii^ ore from the lower Lake ports, it is necessary to carry a wint^s supply piled upon the ground. Such supply for the furnace is picked up by grabs and transferred to feed-bms. At the custom copper and silver-lead smelting works of the Western United States a supply is intended to last from two to six weeks and this locks up much capital while the ore is in process of treatment. Even at smelting plants, treating their own ores, it pays to use from large ore beds which have been bo stored as to be quite regular in composition. Mills turn to have in their feed-bins enough to cany them on overnight, the coarse crushing being done on the day shift only, or in caee of a breakdown, at the supply end.
Ore Bins or Pockets. — These may be flat-bottomed or inclined-bottom bins and may be made of wood or of steel. When they are flat the dischai^ point is at the bottom and side of the bins, then when dischai^ied about half of the contents, forming a natural slope, remains. When the bottom slopes three ways to the side chute most of the ore runs out. This kind of bin is shown in Figs. ISA, 40, 45 and 98. Sometimes the bottom slopes four ways, forming an inverted pyramid. Then, the slopes being steeper, the discharge is better. At times the bin may be of steel, tall cylinders on end with a side-opening at the bott«mi, and the ore forming its natural slope.
Feeders. — There are feeders of many types calculated to give a speedy delivery as into cars or upon a conveyor; or a regular or evrai supply to a furnace, to a crushing nmchine or to a vat: They include:
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fiW ORE STORAGE AND StJPPLY
Ore Bin Gates. — The chute of this gate. Fig. 318, is a continuation of the sloping bottom of the bin. In operation the attendant opens the gate according to the supply he needs, whether to quickly fill a car or to supply a crusher. He must wateh for a sudden rush of ore, or, when this hangs up, be uses a bar to loosen the ore and to make it run. Often there are men stationed above who poke down Mx ore in case &b bin is to be com^^tely emptied.
Shaking Screen and Feeder. — In
Fig. 318.— Single Back Gftt«. Fic. 319.— Combination Shaking-screen
and Feeder.
Fig. 134, at a, is shown a grizzly, Fig. 54, to remove the fines while the lump (oe is crushed. A better separation and a regular feed is obtained by substituting for this the eccentric driven screen feeder, F^. 319.
Fig. 320. — Moving Feeders.
Rotary Feeder. — In Fig. 320 (o) is shown a side elevation of a feeder disk in which as in the shaking screen there is a separation into fine and coarse, the latter passing on to the coarse crusher, while the
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Feeders 565
fine drops between the slowly revolving disks. It uses aminimtun of space horizontally.
Traveling Apioa Feeder. — This, aa shown in Kg. 320 (6), takes a mixed feed whose amount is regulated by a slide at the front of the bin-opening. It effects the same separation of the ore according to fineness. The fine ore is caught in a by-pass chute, which carries it to clear the lower chain. Often a tight apron-feeder is used, delivering all to a car, a conveyor, or to a crusher. Where the ore is fine, or wet or sticky, or contains large lumps, making it liable to bridge or hang up, then steeper slopes than the usual 45° to one ^de of the bin are preferred. Thus a bin of wood may have an inverted pyramid bottom, or in a steel bin a hopper of inverted conical form. This brin^ its discharge at a middle point of the ore column, and so gjves a surer run of the ore. To take this discharge the feeder is well adapted. The bottom opening can be lengthened out
Fio. 321. — Reciprocating Plato-feeder. Fig. 322. — Hanuner Feeder.
liberally above the apron. ' It is also well suited to coarsely crushed ore, especially for tall, narrow feed-hoppers that need a well-assured feed. Such hoppers used as roasters and receiving a wet sticky concentrate such as is produced in flotation, are especially liable to hang up. We may note that this type of feeder cuts down head room to a minimum.
Reciprocating Plate Feeder. — ^This is given in Fig, 321. It receives a quick reciprocating motion from an eccentric and delivers through a wide front opening, the flow being regulated by swing hammers, Fig, 322, set to allow the passage of any large lump then to fall back in place again. Thus large lumpe cannot jam the opening.
Removal of Waste Wood and Tramp-iron from tlie Feed. — In mining the ore chips, wedges, and wood fragments, as well as nails, chains, nuts, and bolts, called tramp iron, are to be found in the ore and prov^on must be made to eliminate these objects as speedily as possible, especially the iron, which, if large enough, may stall or break the crushing machines that follow.
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Oto Ore Storage And Supply
In 8Bin|ding ore the attendant, as already stated, r^plates the flow of ore fnun the feed bins. At the same time he removes the waste wood and biunp mm. The former is less dangerous to the coaise crushing machines, so that but i or i^p of the whole is passed on for finer crushing. Fig. 323 shows how the iron is removed. The ore is earned frcnn the feed bins by conveying belt to deliver to the feed hopper of the crushing machine. The bead pulley of the conveyor is magnetiied and attracting the iron removes it from the ore stream to drop from it as the bejt conveys it from the magnetic field.
The wood tiiat passes
through tiie machines is
cround small and at the
on the water to be removed by hand from time to time.
PUMPS AlfD BLEVATOKS
For conveying concentrate contunlng from 6 to 10 per cent moisture, belt and push conveyors are used, while for elevating purposes belt-bucket elevators are common. Elevators are a source of trouble, and if posable should be avoided.
A variety of different pumps is ou the market for elevating pulp, but those that are really good are few. Machines used for this purpose are the Frenier spiral, centrifugal, three-throw plunger, tur-lift, bucket^levator, and tailing-wheel.
The Frenier is satisfactory for low lifts up to 8 to 10 ft.
A three-throw plunger pump will giv^ trouble with sandy material. The valves cut out quickly and the packing requires renewing often. Airlifts are simple, but require large quantities of air for ordinary lifts of 8 to 10 ft.
The old tailing-wheei gives the least trouble of all. It is reliaUe and the cost of repairs ia low. The one objection against it is the hi^ first cost to install.
For moving sand a well-designed centrifugal pump with white caetiron liners, easily accessible for replacing worn-out parts, will give good satisfaction; the Byron-Jackson pump is an example.
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Centrifugal Pumps
For the movement of sand and fllime pulps, water and solutions, both low-and hi^-preasure pumps are used. For presBure filters and high lifts the three-throw plimger pump is much employed and the centrifugal pump for large vohime.
The Centrtfagal Pump. — Fig. 325 is a view of this tjrpe of pump, taking
its suction at the front and delivering to the down-tumed dischaige pipe at the ri^t end. Having no valves it works well for pumping sand or slime pumps. Fig. 324 shows its internal construction. The rapidly revolving impeUer throws the water to the exterior of the castng, forcing it out of the'casing, when it escapes to the down-turned discharge exit.
$olutioo Puoqw. — Centrifugal, three-throw plunger, and air-lift pumps ate in common use for ele- ^ vating B<4utionB.
Centrifugals are probably used more than any other for elevating to heights of 10 and up to 50 ft. They require considerable attention owing to their high speed.
A three-throw plunger will pump solution to a height of 100 ft. or more. They require little attention, and are economical in every way. I prefer them to centrifugal pumps.
Fig. 326. — Jackaon Centrifugal Pump.
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The air-lift is suitable for lifts not exceeding 10 ft., and is particularly useful about a plant using the counter-current decantatkin process where the lift would not exceed 2 to 3i ft.
The Ftenier No. 1 Sand Pump. — Fig . 326 shows a section of the pump.
Fio. 326. — Section of Frenicr Pump. Fio. 327. — FVeaier Pump ArraQgemeDt.
The trunk or body of the pump, 44 in. diameter, mounted on a horizontal shaft, constitutes a spiral rectangular tube. There are no valves, but the sand and water scooped up at each revolution of the spiral and by the hydrostaUc head created by the revolution flows to the center of the pump, and discharges under pressure up the discharge pipe at the right. The spiral passage with an opening 2} by 6 in. and at 20 R.P.M. will lift 3000 gal. per hour to the height of 14 ft. Due to its simplicity and ease of repair this pump is much liked. Where it is desired to increase the height of dehvery this may be done by introducing an air jet into the discharge pipe as shown in Fig. 327.
The Three-throw Hunger Pun^. — The phinger acts on the down stroke only to press out the pulp or the solution, and „ _ „
the gnt cannot get past it to cut the pj^ '^ ^ "™
cylinder. Having three cylinders there
are three even impulses per revolution. An air chamber at the diachaige ude as seen in the figure also tends to equalize the flow, Fig. 328.
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Chapter Xlvii Cost Of Plant Amd Equipment
Cost Of Plant
Based upon figures of 1913 to ldl5, when prices were comparatively stable, we give data that may serve to indicate the costs of metallurgical plants. We may safely calculate, however, that the costs in 1920 will be double these, but that, when the present abnormal labor costs ag^n return to the older figures, then plant costs will be correspondingly decreased.*
In earlier times, when not so much was done automatically, when wood instead of steel buildings prevailed, and where bedding was done on floors instead of in overhead bins, the cost would have been half that just given. There are drawbacks to much permanent construction in which, where changes are to be made (and this is often what should be done), such changes are expensve as compared with those in l^hter construction.
The first step in such construction is to obtain the services of a competent constructing engineer experienced in the planning and building of the kind of works contemplated. Such service is particularly valuable in the avoiding of expensive alterations, and may amount to 3 to 5 per cent of the total costs. It is so much easier to make changes in the plans than to later correct them in the works themselves. The deare " to make the dirt fly " should be overcome.
Having matured the plans in detail and made estimates of costs, based upon the unit costs as below given, one can obtain bids from manufacturers of machinery and dealers in supplies and is then in position to proceed with actual construction.
To this end tmloading facilities should be provided, a good road and if possible the railroad tracks brought to the site of the works. Ample room should be provided for lumber and for piling it so as to show just where it ia to go. Also suitable storerooms and workshops for the use of the mechanics are to be built. Roomy framing plots and handy places
* It was thou^t that, as after the Civil War of 1861 to 1S65, prices would go down in the coime of two yeara, but we must recollect th&t to-day European nationa are too short of capital and too exhausted by the recent strife to be able to dump goods upon OUT shores in large quantity, so in the United States laboi>demand bids fair to keep up. When priow of labor and suppUes go down then we may expect an increase in
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670 Costs Of Plant And Equipment
for the Btorage of machineTy and its protection against duoage and nist aic also to be arranged for. A supply of water must be available and often one may get a supply of electricity for power and lighting also.
The labor supply must be studied and provision made for the oomScat and efficiency of the men. Without such consideration the building ci & mill in an out-of-the-way place would prove disastrous to an enterprise.
Prdlminaiy W(»k. — ^The cost of all this preliminary work will amount to 5 to 10 per cent ot the total and may be estimated on Mk ground.
For concrete work a quany may have to be opened and a bed of suit' able sand may be available. Indeed it may be prudent to file upon a claim covering their location.
On page 547 we have already discuaaed the nature ai iba site to be chosen and the tights to which the company is entitled in filing upon ground for a proposed mill or smelter site in connection with a mine.
Caipeator work with a picked-up local crew will average S28 to $31 per thousand for framing and erecting, $19 per thousand board feet for siding and roofing, S2.50 a thousand shingles for shingling, and S1.25 per square of 100 sq. ft. for putting on corrugated iron. The niuls needed in erecting would be 18 to 21 lb. per 1000 board feet, in putUng on siding and laying !HQ-flooring; while for 1-in. flooring 28 to 32 lb. per 1000 board feet is needed.
Minor Items ure Important — ^Thus oonsderable lumber is needed ior forms and for staging. The building should be painted, fire prptection and heating arranged for, office and laboratory equipment bough(.
Alteratioiis. — Upon a completed mill it may be necessary to m^f alterations and this, as experience shows, may amount to 5 to 16 per cent of the total costs.
Winter work in the Northern United States or in the mountains may add as much as 33 per cent to the total labor costs even in a mild winter, and in cold mowy weather such costs may rise to 50 per cent. Concrete work often costs 35 per c«nt more, as complete arrangements must be made for heating and protecting ^^^inst the frost until after the preliminary set, after which freeeii^ need not affect it.
E]q>eD8e oi Rebuilding Old Works. — As in a new mill, the costs can be rather accurately figured, but the amount of hardware and lumber that can be used agun is often misleading. The costs of the carpenter work and of the reassembled machinery will generally be twice that of a new plant.
Underestimates. — These are due to guess-work, lack of good or^niiation, omissions and changes in plans, neglect of prelindnaty work, too much reliance placed on general figures, and inefficiency of labor due to unfavorable conditions. Also must be mentioned the danger of strikes, bad weather delays, and failure of railroads or supply houses to supply material as needed.
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Costs Of Plants
Machineiy Prices. — ^A reputable machinery house will give valuable infonnation, and no matter how confident the constructing engineer he should give careful attention to it. They are willing to go into details with him. They do not drop their responsibility when their machinery is delivered and arc always desirous of protecting themaelvee in this way.
Untried innovations, especially by a small plant, should be avoided. let it be tried out by a larger operating company and, if it is fully proved there, it can be put in.
If the plans of the works are carried out, a good oi^animation maintained and efiBcient labor obtained and kept, then the ^ures for construction will be found a little hif^r than actual costs.
COSTS OF HBTALLURGICAL HJUfTS
The following table ^ves the daily capacity and the total cost of a variety of plants based on figures of 1913 and before. Costs are now (1920) easUy double these:
C*p«ltr in Twantr-Four Boun
Iron UaBt-fumace
Add BsBBemer with four remelting
cupolas and hot metal muter
Acid open-hearth ; ten SO-ton fumacea BiMie open he&rth; ten SO-ton fumacce Copper Bmelting and converting
Silver-lead amelting
Refinery for base-bullion
Refinery for dord bata
Zinc smdteiy
Copper blaat-fumace works
Copper bhaUumaoe w<»ks with converter plant
Copper blast'.fumaces and oonverting (Wuboe WoAa, Anacmtda)
0 tons pig iron . .
1000 tons of ore aiudted to 100 tons of 46 per cent matte and this converted to bliater copper
SOD tons mixed lead over to base-buUion
100 tons baB»-biillion, refined by the Parkes jwoceaa
100 tons copper from blister-copper
0 tons 10,680,000
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Costs Of Plant And Equipment
TTUTT COHSTRQCnOH COSTS IN 1014
These are the most useful to the engineer, since, having made plans of & plant or of any proposed building, he can use these units in nmlHug his estimates of costs. These data, found also in engineering hand books, are carefully set forth for a smeltii^ plant in an elaborate paper by E. Horton Jones, Trans. A. I. M. E., XLIX, 3. These figures, quite applicaUe bi the Clifton, Ariz., district in 1914 would need to be doubted to conform to our present 50-cent dollars, and should be modified by any accessible recent costs. At that time common labor cost $2 and skilled labor $4 per day, as compared with something like double that now. It is to be noted, that the figures below given for imit-costs, are averages.
The works coet, completed, $2,105,020.17. Out of this has to be reckoned $100,649.88 for engineering and $140,277.72 for indirect expense, including all necessary for clearing and preparing the site and its approaches, working equipment, personal injxirics, railroad transportation to emploj'ees, etc. This left $1,864,092.47 for the work that would show upon the completion of the plant. The engineering cost was then 5.4 per cent of this, indirect expense 7.53 per cent, a total of 12.93 per cent to be added to the construction costs.
The following is a recapitulation of all costs, with its list of buildings, all equipped:
No.
N.», „*«„....
No.
Nameof AKooiit.
Engineering expeoae. . .
Roaster duat chamber
fiue
Boiler and black-
Receiving bine
smith shop
Cniahing plant
Machine and car-
Sampling plant
penter shop
7H0O
Bedding plant and
General office
bunker bins ...
Roasting plant
Laboratory
Roaster dust chamber
Sample room
Reverbertttory plant.. .
Converter plant
Indirect expense ...
wai
Converter duot chamber
Power-plant
S500
Conveying system
Oil sujJpty Bump and
Reverberatory flue
S2,lO6,O2O.07
The Unit Cost for Concrete Foundations. — The average for all foundation work would be $3.37 for labor and $5.48 for materials, a total of $8.85
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Unit Construction C0Bt8 573
per cubic yard, in place. For reinfoTced foundations this is much moTe expensive, the labor cost being $5.84 and the materials $7.66, or a total of $13.50 per cubic yard in place.
Unit Costs for C<nUTflte Flows. — ^These are laid like a sidewalk 5 to 6 ft. square, 4 to 5 in. thick and with a finished top. They are reckoned at a cost per aquaie foot as foUows:
Plain concrete floors — $0.06 for labor and $0.13 for matMials or a total of $0.21 per Bquare foot. Reinforced concrete floors — $0.18 for labor and S0.23 for material, or in all $0.42 per square foot. These are formed, reinforced, and finished.
Unit CoBt for Excavation. — This depends on its nature, as below given;
For shallow excavation, using wheelbarrows and slips or scrapers, and with a " haul," or distance to move the materials less than 100 ft., the coet averaged $0.81 per cubic yard reckoned in place. When the haul was greater than 100 ft., needing carte, this cost rose to $0.95. Where the ground was solid, needing some blasting, even with less than a 100-ft. haul, the cost was taken $0.84 to $0.93. Like the preceding when the ground was hard, and the haul (using carte, etc.) avwr 100 ft. the coet became $0.89 to $1.00 per cubic yard.
Averaging all the unit costs for excavation we find it to be $0.79 per cuWc yard.
Unit Cost for Electric Lighting. — ^This is based upon the cost for wiring, and the material for each drop or li|^t used. It is averaged at $4.84 for the labor, $5.85 for the materials, or.a total of $10.69 per drop.
Unit Costs for the Erection of Machinery. — The total cost of the machinery is made up of ite coet f.o.b. at the factory, plus freight to its destination, plus the cost of unloading and erecting. This may be computed at BO much per hundred-weight. The manufacturers will quote wei^te and prices at the factory, the freight rates may be obtained from the railroad schedules, and the erection costs are as here presented. It is useful to reckon prices per hundred-weight from the machinery coete as collated, according to the nature of the machinery. The freight coste vary with the classification.
The unit costs, both of the value delivered at the plant, and for erection, vary with their nature.
Group 1 refers to en^e machinery that needs to be placed, cleaned, adjusted, and lined up. The cost of the machines delivered at the plant is $12.67 per himdred-weight or 12.67 cents per pound. Add to this $0.92 for luiloading and erecting, and we have a total cost for the machines in place $13.59 per hundred-weight.
Group 2 is similar to Group 1, but not so heavy and takes proportionately more labor to put in working order. The cost delivered is $8.53, and for erecting $1.50, a total of $10.03 per hundred-wei^t
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574 COerS OF PLANT AND EQUIPMENT
Group 3 u lie&vy machinery needing little labor in erectii^. Tlw machinery delivered coet S1.04, and for erecting tO-68, making a total of f 12.72 per himdred-weight installed.
Group 4. This leeembles Group 3, except that it is electrical. Its cost at the works is S12.67, for erection $1.63, or in all $14.30 per hundredweight.
In all these cases the erection cost is made up of labor and the needed small supphes, as cotton waste, oil, small tools, etc.
Unit Cost of Hasottxy. — This is given for a retaining wall at $6.19 per cuIhg yard.
Unit Costs for Painting. — For painting concrete the labor will average $0.08 and the paint S0.12, or a total of 'tO.20 per square yard for two coats of paint. For painting iron, the oonceponding items would be for labor 10.10 and for materials $0.15, or in all 90.25 per square yard for two coats. Woodworic is cheaper, being but tO.lO per yard in two coats, while painting sash and doors ia expenave, being $0.96 per saah, one door being reckoned as two sashes, and all being three-coat work.
Unit Costi at Roofing. — When the roofing conmste of 1^. sheathing coveted with asbestos, but not painted, the cost was per square of 100 sq. ft., for Ubor $4.06, and f(H- materials $12.40, or in ^ $16.46 laid. Much of the roofing was of 2^. stuff with asbestos covering, takin^; a total cost of $26.08 per square. The costs vary greatly according to the kind of roofing needed.
Unit Costs of Shafting, Pulleys and Belting.— The baas is per hnear foot of shafting equipped with its average of pulleys and belting. According to the building in which it is used it varies from $22.76 in the sampling plant, to aa little as $7.98 per foot in the blacksmith shop.
Unit Costs for Structural SteeL — ^This has reference to the steel used in the construction of trestles, buildings, etc. Like machinery, it includes the cost laid down at the works, plus the cost of erection. It is a pretty uniform figure and will average $87.13 per ton or 4.356 cents per pound.
Unit Costs for VentUatora, Windows and Dochb, Woodwork and Wooden Floors. — Ventilators cost on an avenige $96.67 each, erected; windows and doors were reckoned at $0.81 per square foot. Woodwork cost $52.55 per 1000 sq. ft. board measure, while the wood floors cost $0.21 per square foot erected.
Unit Costs for Labor.— Upon these, b& shown above, other costs depend and it might be well for a rough approximation to vary the costs here shown on the basis of the laboi^«ost. Thus, with the cost of labor doubled, we may expect that supplies have correspondii^y increased, and so that the above estimates are to be doubled. The proper way is, however, to readjust the labor costs as given, by the new figures, for labor, and ascer* tain freshly the cost for supplies and equipment.
Composite Costs 575
Wage Seals.— 'niifl wu in September, 1913, at CUftoo, Aris., for common labor S2, for skilled labor $4 per day of ei^t houn. We may note that of late, due to increasing wagea, and the scarcity of labor, men have become less efficient.
coicposnx COSTS
These are convenient figures for arriving at an approximate idea of the cost of a building, either empty or equipped, based upon ita area or cubic contente cm* of its machinery and equiisnent, according to its capacity.
The cost of buildings varieB from as little as SI. 51 for the roasting plant, to as much as S3.62 for the crushing plant per square fooi of floor .atsa. In re erence to cubic contents the cost varied fTom 40.11 to S0.22 per cubic foot for the respective buildings just quoted.
When it comes to the cost of these buildings with aQ tb^ machinery or equipment in place, the figures arc increased according to tiie cost of that equipment. Thus the roasting plant comes to $4.76 fuid the crushing plant to $5.62 per square foot of floor area, while, for cuImc contents, the former is reckoned at S0.33 and the latter S0.34 per cubic foot, showing how expensve relatively is the equipment of the roaster-plant.
The feed-bins asid the bedding-floors (using the Messter System) cost tO.66 per cubic foot of capacity, the heavy receiving-bins oome to $3.M for the same unit.
Conveyors will cost from $19.02 to $32.68 per ton of houriy capacity, this including the cost of the steel supporting structure. If we were reckoning a conveyor according to its cost per linear foot, erected, this would be $34.47 per foot.
Dust-chambers are $0.30 on an average per culnc foot, the flues $0.45.
A useful figure is, that the cost for power-house instellation (including the boilers), would be $55.32 per indicated horsepower, while, if the boiler plant is not included, this drops to $37.40 for the same unit.
Again, the cost per ton of output in twenty-four hours for the leverberatories is $43.47 per ton.
The cost per roaster, complete wit^ it« instellation, ia $17,091.86, tmd, with Ub building tmd flues and dust-chambers, this figure rises to $24,097.34.
The railroad trackage cost $4.64 per running foot, being $2.71 for labor and $1.94 for materials.
Raw-material Prices. — In Mr. Jcmes' paper a tet^ of audi supplies is given f .o.b. Clifton, during 1913, and to this the student is referred.
The author is of the opinion that a careful study of this paper, including ite illustrated details and descriptions, would constitute a great aid in an engineering education.
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Part Ix The Business Of Metallurgy
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CHAPTER XLVIII THE GENERAL ECONOinC SITDATIOH
DISTRIBtrnOH OF WEALTH
So far as tiie United States, in ita internal economic condition is concemed, we may say:
The fixed wealth of the United States in 1916 was about $260,000,000,000, whereof about $30,000,000,000 was in stocks of goods and all the rest in real estate, railways, etc. The population of the (Country was about 102,- 500,000 souls, of whom about 41,000,000, men and women, were workers, about 14,000,000 of them being farmers. The total national produce was about $1,200,000,000 Urns of goods, worth about $45,000,000,000 to $50,000,000,000. Out of that produce a group aggregating a little more than 400,000, who received incomes in excess of $3000 and paid income taxes, got about $7,900,000,000. Less than one-half of that was derived from inveatmenta and more than one-half came from U»e personal efforts of this class. Persons enjoying incomes of less than $3000 received about 44 per cent of the dividends paid by corporations and a much larger proportion, perhaps 75 per cent of the government, state, municipal and corporate interest payments, lliere renuuned from $23,000- 000,000 to ^8,000,000,000 to be divided among 27,000,000 non-agricultural workers, who received an average of somewhere betweeen $855 and $1040 each. Among the great classes of workers there is a wide difference ia earnings. The farm hand in 1916 averaged about $400, the factory worker $675, the steam railway man $886, and the metal miner $1260. Some claases probably averted higher wages than the metal miner.
A satisfactory economic system can be based only on natural hmnan impulses, and of these the moet fundamental is self-interest. Increased production is at the present moment the most presang national need, but it will become effective only when for every man increased production becomes the talisman by which bis paper wages can be turned to gold.
ECOnomCS of EnGINEERING
The mining engineer, entering upon the practice of his profesaon, may confine himself to the technique of mine-operating, while the ore, delivered from under^round, is then taken in charge by the metallurgical engineer.
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fiSO THE GENERAL ECONOMIC SITUATION
whose buaiiiess it is to win the metals from it. The young engineer, entering metallurgical practice, takes subordinate work, such as drafting or assaying or testing, which gives him thorough knowledge of certun branches of the work that he is to take up later in operating. On the other hand, the duties of metallut^cal practice may be barly assigned to rather the draftsman or assayer, so that they may be compelled to think al<nig the lines of actiial practice. We find, as a matter of fact, Uiat men actively operating, are thinking much of those duties, are studying and discussiiig them, often relegating to the background the economic conaderations later liable to come up; hence the discusmim given on the following pages to these aspects of metallurgical engineering.
The Ectmomic Situation in the United States as Related to the Prodnctipn of Metals. — The prosperity of custom works re6ects that of rniTiing and profits in them fall away in dull times, since their charges must be reasonable in order to get the ores. With the mines plants it is different and their prosperity is tied up with that of the mine. Milling or smelting the ore is but one item in mines operation.
The value of a metal is fixed by the cost of production at the " marginal class of mines," that is those mines that just pay their way. If the price of the metal goes up, leaner mines may become marguial, while the firel ones cited come into the profitable class. If the price of the metal drops the marginal mine must close down.
Thb Labor Situation
As the country has opened up so has the mining industiy, and the demand for labor has been met in part by European immigration. This has of late practically ceased so that workingmen are acutely needed; and wages have doubled. Organized labor has taken advantage of this to make demands under pretense of needing a " living wage," that is money enough to meet the necessities, but also many of the luxuries of life. Often the strikers have not cared whether they worked or not; a hohday would well suit them. As a result, capital, whose rewards depend upon uninterrupted operatitm, has lost seriously, and the marginal mines are having to shut down, thiis again cutting down the supply to custom plaits.
nnioos and Non-union Labor. — ^Two methods are in use in industrial plants, viz., the open shop and the unionized shop. In the case of the open shop the method of individual bargaining prevails; in the second, that of collective bargaining, that is, the agreement for wages, hours, and treatment, are made by the officers, or by a committee on bel&lf <rf the men, who belong to a labor union. In the second method it is expected that none but union men are to be employed.
The union shop with its working force is controlled by tbe labor union.
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The Labor Situation 581
Workii^pneii who treat individually, lacking the backing of a union, may be taken advantage of by the employer, who ofiers them a low wage or treats them in an arbitrary way, and such men wish the support of the union. The union declares the equality of all their men, says that the fast workman shall do no more than the slow one, because the fast man compels the slower one to work to exhaxistion; that, if output is increased, then demand will cease and the workman will be out of a job.
The worst feature of a union is, however, its tyrannical power, that it makes its demands on the penalty of a strike to enforce them, that the power to call a strike is entrusted to intermediary ofScers, when it is notorious that such positions have been sometimes g^ed by cajolery, bribery, and the methods of ward politicians. In such cases strikes have been called when but a small fraction of the working force has desii^ them. In a unionized works men who are employed in special work may be unrepresented in committee. There may be a dozen or more of such specialized positions enjoying compensation dependent on their skill. There will always be an incentive on the part of the committee man to favor his own job, or his own friends, and on the other hand the works manager may be only too willing to back the committee man if he sees it is to his advantage. Such methods produce discontent, and eventually a strike. Even men who receive the hi^est pay may be so affected. It is suggested that these highly paid men might be paid even more in order to keep them quiet, but there will come a time when this is more than the union as a whole will stand, since the action of the committee is not final. The signature of the company bears with it responsibility, but the signatures of the committee do not. The union is not incorporated; it has no tangible assets, it is irresponsible. It cannot bind an individual to work, and if there is a good demand for labor he may seek it elsewhere. If indeed the union wishes to aid some other it may go out on a sympathetic strike, due to no fault of the responsible company.
Where large bodies of skilled men of one trade jtnn in a union, that is different, but for a works having varying pay according to the skill of the workmui the union is bound to be inequitable. The total unionized labor in ^e United States is 3,000,000 out of 35,000,000 people engaged m gainful occupations; it constitutes a body one-twelfth of all the workers, holding up production if striking, and raising the cost of living to all the rest, and especially the many income receivers, who, due to fjieir thrift, have invested in these very companies. Thus a man who has thought to provide for his retirement in old age, or to insure on behalf of his loved ones, is compelled to labor on, or to live a constricted existence.
The labor union also makes the tyrannical demand that non-union men shall not work with union men, that the shop ^lall be a " closed " one, even that goods made t^ non-imion men shall not be delivered to a
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882 The General Economic Situation
unirai t^op. Many modem strikes are based on theae ideas, and the strikers are prepared to cany them out with picketing or even with deadly violence.
The union disapproves of labor-fiaving machines, unless the profits arising from their use are distributed among all the men, its ressrai being that it fears production will outrun demand and so liie workmen wiU ha>'e nothing to do. A sufficient reason why such machines are not contrary to the interest of the workmen lies in that fact that formerly, when such work was done by hand, a skilled man had to be physically superior and by middle life, despite his skilled knowledge, had to take inferior work. With the introduction of the machine he could retain his employment^ indeed earn more than when he laboriously worked. It was of mutual advantage to retain the services of such an experienced and trusted man.
This implies the settlement of disputes between employer and emi^oyee. By enactment of law it may be compulsory or advisory. By private agreement an arbitration committee depends on its power of persuasion, or the willingness of both sides to submit. In this, public opinion may come in as a factor, especially if its interests are directly involved. In so-called compulsory arbitration one can hardly see how a works can be compelled to operate at a loss except by confiscation, nor how a man, or even many of them, can be compelled to work if they do not wish to; they are pecuniarily irresponsible. To be sure, as has often been done, when the eanployer has submitted, it has been possible for him to raise his prices to meet the increased labor^ost, and so pass them on to the ultimate consumer. Art»- trating committees may be appointed to adji^t grievances, but a better way is that each man who feels he has been ill-treated should have access to the works-manager or superintendent, the ultimate judge. In this way favoritiun or nepotism may soon become imknown, and injustice checked in its beginning. Investigation shows in general that the man*s statements are correct, and a fair and eqxiitable array^ement should be made. In these cases it is well that the matter be discussed in private and that, for psychological reasons, the complainant in discuasiog it should eit down.
Where work is abundant and jobs waiting, men become arbitrary, notional and imreasoning, since they feel they can compel. When times are dull and work hard to find, they will take what they can get; the employer may then become arbitrary. We give herewith the periods of panics and their causes which can be used in guiding future action on the part, at least, of the employer.
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Association Of Employers 583
Financial Crises In The United States
Seoonrf 1837
Third 1857
Fourth 1873
Seventh 1907
Eighth — BcvMe decline in faufliaeeB 1913
Ninth T 1927 or 1926
Association Of Employers
We have mentioned the shortcomingB of the workingman; the question
arises: What is the employer to do? His best plan is to imitate the
methods of the union. Association: This is all-important. Thus the
mining industry, the smelting industry, the milling industry should unite to :
(1) Fonn a union to which all should liberally contribute, probably small and large alike.
(2) Appoint an executive committee to establish a propaganda based upon exact and statistical information. These data should be accessible to writers who can make good use of tbem and these writers should be properly compensated. There should be statisticians who have power to enter into the matters of costs and prices, also men skilled in deducing conclusions from them. Income-receiving men of limited means should be hired to preach this propaganda, and carefully instructed. Where the atuation is fitting the officials of the companies should speak at meetings and elsewhere. The association should tise the lockout judiciously and should carefully prepare for the shut-down.
Finally, even as the union man will strike and starve to gain his ends, 80 must the company do, confident that by lockouts a permanent improvement can be made. Where one or two out of many shut down, it is but fiur they should draw compensation from the general fund.
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Chapter Xlix
Organization And Operating
OROAinZATION OF A HBTALLURGICAL COHPAITT
Metallurgical operations on a commercial scale require, generally, the organization of a company, or if the company is already organized, the establishment of a department to provide the additional function.
Where a metaUurgical company is to he organized, the promotets or organizers obtain a charter, or articles of incorporation, from the State in which they demre to incorporate. They next hold a meeting at which they receive the property that is to be taken over by the company, adopt a set of l^-law8 for the guidance of the company, and elect the directors that aie to manage the affairs. The directors proceed to the election <A the cor~ porate officers of the company from their number. The officers of a small company are the president, the yiee-premdent, the secretary and the treasurer. The directors may appoint from their mmiber a managing director, or they appoint a manager from the outside to have charge of the affairs of the company.
THE ADHUnSTRATIVE DEPARTUBHTr
In outlining the organization of a company undertaking metallurgical works, the manager should be guided by the following rules:
He should see that a supreme authority is provided over all action to be taken, and should carefully and fully outline the authority and respondbility of each position, making the duties of each conform to the capability of the party holding it. To do this, he must avoid making any person subordinate to two or more, should place the authority and respcmsibUity together; should distribute the work and the duties not to overburden nor to underload; and should arrange the positions so that pro* motion can come from them. While the manager gives his chief attention to the commercial or business affairs of the company, he generally appcauts a superintendent to attend to the technical affairs of the plant.
The industrial organization, under charge of a manager, would include (1) The operating department; (2) the accounting department, and (3) the purchasing and selling and supply department. 584
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Operating Department 585
(1) The operatiiig department haa to do with all that pertains to the reduction or manufacture of the ore into metal (the winning of the metal from ore) or to refining metals to bring them into marketable form, and has control of the operating forces, consiBting of the foremen (and men under them), the repair force (consisting of mechanics and their helpers, who keep the plant in repair and put in the needed improvements), and the laboratory or assay-office force.
(2) Ttw accounting department attends to the accounting, pay-roll, costrkeeping, and the distribution of costa.
(3) The purchasing, selling and simply dqtaitment attends to the pui^ chase of ore, fuel, fluxes, and the chemical and other supplies. It sells the products of the works. By-products, in process of farther treatment, are not included.
THE OraHATIHG DEPARTHEHT
We discuss the qTiahfications of those in charge, the management of the working force, their welfare, efficiency, and their payment.
Duties <d the Superintendent — The superintendent not only must be informed as to tiie actxial technical operations, but he must know how to organize his force. He should be able to handle men effectively through tact, discretion, and firmness. He should be strict but just; able to encourage as weU as to drive. He is often the metallurgist and constructing engineer as well as the superintendent; and has the direct management, with the aid of his assistants and foremen, of the furnaces and the metallui^cal machinery.
When things go wrong he may be called on at any hour to correct them; if a furnace is in bad condition, or a machine out of order, he is responsible. When all is going smoothly bis duties may be li^t, but when troubles come, or the company is losing money, his work is hard. If he fuls in adjusting difficulties, no excuse is accepted; he must succeed or res^. Much of his success depends on his subordinates, and first in importance among tbem the foremen.
General orders, applying to different departments, shotUd be issued in multiple BO that each foreman affected shall have a copy, thus avoiding delay or misunderstanding.
The work of supervision and control, that is, for the superintendent, his assistants, the foremen, the testing and laboratory force, and for the offiee men is paid monthly.
The men chosen to take charge of the different departments should be chosen according to their qualifications for that particular work, essential qUahties being intelligence and reliability. Lower-^rade labor is used for plain, hard, routine work, the better labor where judgment is neceseary. Common or unskilled labor is liable to make blunders; however, when
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586 Organization And Operating
trained, if faiUiful, it becomes reliable. Tbe saving made by the emi^oy-ment of cheap men for operating coetly machines is offset by the loss trf time, or by actual disaster.
Duties d ForemuL — ^A foreman is often one who has been advanced from a lower position in tbe same works, or be may have been selected from another establishment. With a view to efficiency it is well in selectr ing a foreman to find out from former employers bow he gets on with his men, whether he finds fault with them, whether he has been threats ened by them. Thouf^ he may lead off or show them how thing? are best performed, in general, he has enough to do in seeing that the work is well planned, and that the working force are busy. He has not only to note tlie execution of tbe work, but to plan ahead, to be sure that everything is provided for, and is at hand when needed. In much of such woric he need not drive as in routine work.
Testing or Research Work. — This department needs a head, not only capable of carefully making teats, but also of drawing useful deductaons from them. Thus, a test may corndst in determining which of two or more methods will be tbe most effective or economical. For example, one may wish to lean what coke (taking account of its price) would cost the least per ton of chai^, or which of several methods of admitting ur to <Hie or other of the hearths of a MacDoi^all roaster will give tbe best roast.
Tlie chemist and assayer are called on for the results of tbe analysis (rf by-product0 of tbe works, of the ores purchased, and of the products sold. He must produce with pnnnptness results that control operations. In the case of oi«s bought, and of products sold, accuracy is fundamental. Certain supplies like oils and chemicals may have to be analyzed by him.
Repair Force. — The repair force, consisting of the masteiMnechanJc and skilled men under him, not only have to make repairs, but have nev construction to attend to, generally under supervision of the supeiintendent, who may, where the works i^uiic, employ a constructing en^neer and draftsmen. It is a good rule, in case of a breakdown, or other amilar emei^ency, that this work have precedence, and other work be dropped to expedite it. A plant might easily be loedi^ a dollar per minute during such a period.
Wotkxooa. — The laborers at a smelting plant are largely unddUed. These are called outnde men or " roustabouts." They do the worit needing pick and shovel, such as imloadir^ etas, handling the products of the works, carrying materials, etc. Skilled laborers, or men working in shifts of eight hours daily, and called inside men, receive higher pay per day than common laborers, the pay varying acording to the particular position hdd. These men are responsible for the successful performance of the duties asragned to them.
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Rules Of Wokk
Rules Of Work
For keeping discipline, and to prevent slackDess, certain rules, the result of long experience, have been laid down. These are:
The men miiat be promptly at work and must work full time.
Inside men must be on hand the entire time of their shift, and must eat their limcheon as they can spare the time, while not neglecting their duties. Cbai^e-wheelen must keep up the supply, but may rest at intervals as they wish. They are not called on to do other work, except sweeping up their own places before going off diift. The inside man can leave when relieved by his partner, but must wait for the partner tmtil relieved. If the latter fails to appear the foreman provides anoliier man, who then holds the place, the absent man losing it, unless he has a good excuse, or if ack, he is expected to notify the foreman who provides a man for tlie place. When the absent man desires to return to wcn-k, he must notify the foreman one shift in advance, so that the substitute is not put out of a shift for which he has come prepared.
When men are sick on shift, if not too seriously, they should be held, if possible to the end of the shift. It is impressed on them that it is detrimental to the work for them to leave, and that it is difficult at such short notice to get a substitute.
Men must obey orders, and disobedience can only be followed by discharge, otherwise discipline is weakened.
Let the foreman be strict but just. It helps in discipline to let out a poor man occasionally, and if this is seldom done one may suspect that the foreman is not strict enough.
Do not entrust men to do routine work without supervision and inspection, they may do it wrong or become careless if they realise they are not watched.
In mneltmg or milling, operations are carried on by a crew who work t(^pther, each being responsible for his assigned duty. All are directed by the foreman, who sees that operations are regular. AU this crew an inside men, and a fresh crew replaces the precedii^ one.
PLAHT-OPBRATIOn
Efficiency of Men. — In smelting, mechanical appliances are increasingly in use, demanding a greater investment of ca^utal, thus reducing the labor per ton of ore treated. This makes a man's work less strenuous, and yet better paid. Both these advantages make a man more anxious to letAin his job, and so he is more dependable. In a large works the labor needed per ton of ore is less than in a small one. Steam-power needs more attendance than water-power, and is on the whole more expensive.
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588 Organization And Operating
In Btarting a plant into operation, a list of places and occupatione o^ the men is prepared, bo that men who are chosen may be quickly aseigneu to their positionB. These men should be questioned and carefully chosen. When a new works is about to start, skilled men often apply, and they may be willing to do common labor pending the starting of the works, and in this way be held until their services are needed.
Care of Men. — Provision is made for the care of the men in case of sickness. A charge of $1 per month is often made against every man, and this entitles bim to medical attendance and care at the hoBi»tal in case of accident. If a man works five days in any month, the $1 is deducted from his pay for bosjHtat dues.
MORALS or nrsms hen
Ad efficient mill or smelter man is proud of his work, and to encourage this the plant should be kept clean and orderly on all shifts, even thou^ this adds to the expenses. He should have training in repuring his machines and not have, at least for minor repwis, to wait (or the repair gang-
Where those in chai^ are out of sympathy with their men, and hold alocrf from them, it conduces to a larger turnover, that is to men leaving and others in their places having to be broken into the work, involving loflses in so doing. Where possible, and it generally is, the foreman should notify the superintendent of his intention to discharge a man, thus preventing its depending on impulse or dislike.
A man should be fitted to his work, and if he fails to do well, he may be shifted to another duty, better suited to his capacity or tastes, rather than that he should be let out. Of late, for a lai^ works, an employment-specialist tests applicants to determine not only their fitness but also the kind of job at which they can do their best. A well-trained man should be encouraged to stay. The practice permitting frequent overtime is unsound, since it tends to excite by the prospect of overtime pay, but is, in the long run, exhausting.
It is well to get in the way of talkii^ over works-methods or improvements with the men; it adds to their efficiency and interest in their work and men sugpjsting improvement should be rewarded in some way. The mill man's interest is increased by his receiving training on the repair gang, then he can make minor rep^rs himself.
Contentment is increased by fair wages, bonuses, comfortable and attractive surroundings, yearly vacation, sick pay, and medical attention. Quarters should be arranged for married men, and for single ones bunk houses, where they can sleep undisturbed when off shift. If, however, money is freely spent for these purposes there is a fear that the men may
Modes Of Payment 589
•■■•Jiik the company is rich, so why not strike for higher wages. There is sii^li a thing as being too conciliatory. Even where the company can Ekffiorcl it they ^ould rather wait for the men to ask for the improvements, though these may be put in on the basis of increased resultant efficiency and for the reason that it is the custom in other plants. The men reason thus: These improvements cost much money; we do not value many of them, so why not pay their cost directly to us as wages.
Sunday closing is lecf^nized in principle as ^ving rest and a change in the total life of entire plant force. To interrupt weekly plant operations and to bring back regularity of operation on Monday may involve serious ^cpenses and loss in extraction. Moreover, that men put in this idle time in a manner detrimental to their efficiency is another ai^ument used. Still, the endeavor in any plant is to cut out operations that can be deferred until Monday, such as the work of most of the office force,
Modes Of Paymbht
Daily wages and frequent payments — especially in dealing with laboring ineai of the primitive races. Monthly payments interfere with the steady operation of the works, because at the time of the monthly payment the men are disposed to " lay-off " to spend their money. To overcome the difficulty two methods have been tried. One is that of daily payments, by which a man, who spends his money when he gets it, has only sufficient to supply his daily wants and those of bis family, and none remaining for drunkenness or gEimbling. The other ^^tem is to pay a wage to which is added a premiiun that increases with the time worked. It is paid at the end of the month if the man works through the month, but otiierwise not. This tends to keep him steadily at work.
Also careful attention must be paid to the exactness of the payroll, otherwise dissatisfaction may result, with desertion at a critical time, so that the works suffer from labor shortage. This difficulty is now overcome by the use of indicator clocks by which the time of entry and exit is punched on the employee's card, giving an tmerring record that he can reckon up for himself at the end of the month.
Piece Work. — ^This is the payment prevalent in steel works where the men are paid according to quantity. The great difficulty is to establish the rate which shall be fair for both sides. If the employer sees that the man is making inordinate profits, due to his speeding up, he may cut the rate, to the dissatisfaction or even the loss of the man. The method becomes compBcated where several men work together, since some of them aie more efficient than the rest. A subcontract, however, may be given one man who hires his assistants. In practice, in certain steel works, wages ha/e increased 50 per cent to 60 per cent and production doubled.
590 Organization And Operating
Dafly Wages and Premium. — To aid in this, the plan of pTogreaBi\-e payments has been instituted, i.e., increased payments for tons in exoe^ of normal output. At the Vieille Montague ziuc works several Bystents of payment are adopted, but the mill and smelter men get, in addition to a Sxed wage, a premium calculated <m output, and another premium for tmusual energy. In some cases the men who fire the fumace-blockB get a premium based on the time their furnace laste without repair. In other cases thejretort man gets a premium for all over the calculated percentage of tine yielded. Two-third the premium is paid monthly, the rest is retained to be paid at the end of the year, but only if the man has worked regularly throughout the year. At an English iron works puddlers are paid by the ton with a premium for the ftill nxmiber of ahifta durii^ the week.
Profit Sharing. — ^Tbe system of participation in profits on the pert of the men is both deceptive and dangeroiis. It is possible to admit officials, foremen, and specially skilled workmen to participation, but workmen in general are not fitted for the change. Everything goes well as long as the works are carried along at a profit, but in bad times discontent soon breaks out. The system is not favored by the workmen themselves. They are perfectly willing to share in the profits, but they object to responability for loss, or to even stand for the creation of a reserve fund to cover possible future losses. They cannot aw^t better times, nor can they work their turn at a loss in order to retain their places. Besides this, profits appear too remote, and they cannot understand the relation that exists between the work and the annual profit.
Three methods of profit-sharing have been devised. (1) The workman gets a share in the atmual cash bonus; (2) This bonus is kept back for a specified period and pud him together with the accumulated interest; (3) At the Vieille Montagne works the one-third portion that comes to the steady worker at the end trf the year is this distributed, vii., a portion of it is at once given him, the rest invested on his behalf. The system here has prevented strike.
George W, Perkins, a well-known financier, proposes as a correct profile sharing plan the following:
■{i) That the business shall first of all earn operating expenses, depmdation (a serious item for a mine plant which is a wasting asset) and a fair return of 4 per cent or more according to the stability of the venture on the capital invested.
(2) That all profits above this should be shared on a percentage basis. Thus for a silver-lead plant the labor cost is given at 25 per cent.
(3) That the share of capital should be carried to surplus, the share of labor should be distributed to them as a security (bond or stock), of the company. The employees' share of the profits to be allotted on the basis
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Capital Requibement8 S91
of their pay and that each employee should be reqiiired to bold his oecuiity for from three to five years.
CAPITAL REQUntEHENTS
Tias includes not only capital investment and the funds to meet future oUigations, but also the woi^big or quick aqtttal called the liquid assets, this latter being that needed to operate the plant. It includes the funds needed for purchase of ore, flux, fuels, and supplies, to pay wages and salaries, and to meet incidental working expenses. This includes product bought and in process of treatment. Often it takes thirty to sixty days to get returns oo finished products. It often happens that the buying company or a bank wiU advance money to the seller on his product while awaiting the returns.
Capital Invotred. — ^We have shown that a custom |Jant needs larger capital to cany a stock of ores and other supplies, but with a worica plant the ore is apt to be but a short time to process, and so, especiaUy when it ships a finished product, can pay its way with little trouble. Ca|Htal we understand to mean, not only that invested in plant, but that needed to carry a stock of ore, of supplies as well as to pay wages as they come due. Capital when invested in mines-plaiits rightly expects a large retxim on the investment, because of the risk involved, that the property may not prove up as expected. For we must remember that not only must dividends be lai^, but there must be the return of the ori^al investment. When a mine ceases operation its plant and equipment are but little better than scrap.
It has been the fashion to complun that capital is in the hands of a few greedy rich men. In truth, however, mining stock is well diffused throughout the community and much of it has little value, or in other words the investment has not proved to be profitable. The chance of malting large profits has enticed many into such investments and it has been said with truth that as much money is dropped in these ventures as has been taken out of them.
The Costs of Production. — These may be divided into prime costs, general costs and administration costs. Or we may divide them into working costs and overhead costs. Prime costs are those which vary according to the tonnage put through and which cease if the works shut down. They are also called flat, actual, or direct costs. They are made up of labor, motive power, fuel, material, supplies and repairs. Under material is classed what is used over and over again in specific operations, but which is gradually consumed as zinc in the Farkes process or the tank acid in electrolytic copper refining. Supphes include tools and other incidentals , obtuned from the general storeroom. Repairs inchide not only the labor, hut the parts needed to replace those worn out.
592 Organization And Operating
General Costs.— These include takin^ care of the items of interest, cartages, lighting, foremen, watchmen, miscellaneous labor, sampling, the assay and chemical laboratory, testing, etc.
Administratioa cost takes in salaries, office expenses, law expmaes, advertising, traveling expenses, purchasing, shippii^, selling, taxes, rents, etc.
The Accounting Department
The Council of the Institution of Mining and Metallurgy (British) unanimously adopted a uniform system of accounts, the outline of which wiU be followed as it relates to metallurgical works.
These accounts, while serving for guidance in the distribution cd expenditure and the collection of revenue, must also show the pro6t or loes or the financial condition of the undertaking and should supply the management with the data necessary to check the efficiency of the adminisbation and to pcnnt out how useless expenses may be cut down or revenue increased.
Capital Eqieuditurc. — Before the producing stage of the works is reached, all expenditures should be charged to capital account under the following heads: (1) Lands; (2) Machinery and Plant; (3) Buildings; (4) Surface works as reservoirs, water service, railways, sidings, roads, power line.
After the plant is producing all other expenditure should be carried to " general expenditure account," to be distributed proportionately over the remaining heads of expenditure, but remembering to deduct any levenxie that may have been received. Also, if additional property is purchased, or additional buildings, machinery, plant, or surface works put up, intended to increase production, improve recovery or decrease costs, then these items should be added to capital expenditure and they should bear ih6 share of administration or general charges. If any machinery, plant, or buildings should be entirely superseded or replaced, their cost should be taken frmn cai»tal expenditure and be charged to profit and loss. In case the item is small it may be charged in full, if lai^, then by installments spread over a judicious period.
Ten^toruy or Distribution Accounts. — To distribute costs as evenly as possible it is well to open temporary accounts in order to spread the cost of considerable items of machinery or plant replacement, or payments which occur annually or less often over the monthly costs.
ValuatuMis. — Bullion, concentrates or other marketable products ready for shipment should show as assets on the balance sheet and should be credited to Kevenue Account at net valuation after deducting all realism-tion or marketing expenses. Unfinished products, still in process, ^ould be reckoned at cost, provided such cost is lees than muket value.
Depredation. — This may be considered as it affects capital expendi-
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Accounting Department 593
ture, as already given. Systematic depreciatioD is theoretically correct. It is the amount charged annually to the profit and loea account, according to the conservatively estimated life of each item. The equipment of a mill or smelter plant may depreciate, for example at tlie rate of 10 per cent per annum upon the original capital, the buildings at the rate of 5 per cent. Thus, even if kept in repair, the equipment will, at the end of ten years, have value equivalent to scrap and may be practically obsolete. Depreciation may be divided as follows:
Maintenance, referring to the wear and tear on equipment and buildings. It varies with different classes of equipment, accidents due to deteriorar tion, etc.
Replacement. — Csxised by wear which cannot be repaired without replacement of the worn-out parte. The equipment will therefore fall below its original value.
Obsolete. — Due to new types of improved equipment which are necessary for rapid economical production. The old nuichine can only be scrapped and has no value except as old iron.
Neglect. — Even if properly maintained, the equipment, due to ne^ect, may fall below its actual market or working value.
Inadequate. — ^Machines may become too small to be of service. They often render co-operating equipment of no use, since they are inadequate to serve it. Such machines have a value if sold to thoee whose opemtions need them.
Repair and Haintenance Costs. — Suitable provision should be gradually made out of income for new buildings and equipment. It accumulates a fund for new plant and, at the same time, reduces assets to something like their true value.
Costs. — These may be divided into: Flat or prime costs, otherwise specified as oro treatment or reduction charges, and second, General expense or fixed charges, which take in administration and general charges, realism-tion charges, taxes, and royalties.
To obtain a just idea of costs, where mine and mill are in one, we should really segregate the costs of raining, concentrating, and reduction. The costs we discuss, however, are the metallurgical ones and it is expected that they should be used in connection with the related ones.
Ore TVeatment or Reduction Chaiges. — ^This should include all costs from the time the ore is delivered at the plant until the bullion, metal, or marketable product is obtained, and may be divided to suit circumstances. In it ^ould come (1) power, (2) stores, (3) sampling, (4) assaying, (5) maintenance and repairs, (6) salaries and payroll. Maintenance and repairs should be segregated either on the basis of the shop expenses or according to the labor employed in each department.
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6M ORGANIZATION AND OraSUTINO
ADHraisnuLnoN aud geiteral chargbs
These include (1) oraunilting engineer and graiend mauager'B fees, (2) office staff, (3) stationery, poBtage and telegrams, (4) medical and eanitary expenses, (5) traveling expenses, (6) fire insurance, (7) aaa.- ployers' liability insurance, (8) h«wi1jng, (g) bank charges, (10) auditors' fees, (11) legal expense. (Legal wir^"^ includes questions rotating to tbe rig^tKjf-way for ditches, flumes, roads, and electric power line (see page 549). llie ocanpeny must be defended against those fannen who claim alleged damage done by smi^Ee. Contracts for the purchase of ore and supplies and the sale of products, as well as for railroad frei^t and for railroad facilities should be legally drawn up. The question of royalties and labor and supply contracts are settled Yyy legal advice. (Labor disturbances also need to be legally handled.) (12) Directtns' fees, (14) foreign agency expense, (15) interest on loans and bonds.
Realization Chazget. — (1) Hauling, (2) frci^t, (3) shipping chaises, agency, and commission, (4) sea frei^t, insurance, etc, (5) bujnDg and selling expense, (6) advertising.
Rents, Taxes and RtTitties.— These include (1) nnta on buildings, etc., (2) tax on profits, both regular and excess, (3) public taxes, (4) n^ral-ties.
Reports. — The aimual report gf the manager should show: (1) The quantity ai ore treated in tona with the value per ton, the relative values of recovery per ton in each department; (2) a detailed summary of working costs subdivided to correspond with the main headings fuul subheading <^ the cost sheets; (3) a short tabulated statement d the nature of and expenditure upon new plant and equipment, showing sales ot old plant, if any.
Statements as to tlie quantities and values of supplies, fuel, fluxes, ores, and products on band should be certified by responsible officials, countersigned by the manager. The taking of an independent inventory is advisable from time to time.
Cost Accounts are kept for the purpose of determining accurately the cost oi ore treatment, so as to have a basis for treatment charge; to judge how weU operations are proceeding, and to supply data tar plant efficienqr,
Sfirer-lead or Copier Smdting WoAs. — This may ctHuprise the following responsible men mider charge of the superintendent :
Clerk and metallurgical bookkeeper; testing engjneera in control of the phyacal and chemical laboratory of the research work and tlie test^ ing; engineers and mechanics who have charge of the power plant and of
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Operating Department 596
mscbine repairs; foremen to attend to the sampling, to the yard or outside work, to the roasters, to the sintering machines, if any; and to the blast-furnaces.
Gold Stamp Mill Producing Concentrate. — ^The force may include foremen, mill engineers and repair force, amalgamators, feeders and laborers.
The foreman has general superviaon of the mill and Looks after the handling, cleaning, and retorting of all amalgam collected. The amalgamators drees the chuck-blocks and plates, and keep them in good condition. They set tappets, regulate the water-supply, and make renewals. The feeders attend to the uniform feeding of the batteries, and assist the amalgamators in renewals and at the clean-up. A good feeder is a valuable man about a mill. The vannennen attend to the vanners, or concentrating tables. They must be men with experience, and conamonly should first serve at the vanner as " sulphide-puUers." The crushed^men feed the crushers with the mine-ore as it comes to the mill. Oilers oil the machinery. Sulphide-pullers remove the concentrate or sulphide from the vanner boxes, and store it for shipment. Engineers run the power-plant, and have charge of the firemen. Firemen fire the boilers and remove the ashes. Coal-passers wheel in coal from the coal-pile to the boilers.
On repairs there are carpenters, with laborers to help them. In repairs on the vanners there is a special vannerman to assist.
Porty-stamp Silver MilL — The inside labor may be given as 6 panmen, 3 helpers, and 15 tankmen on eight^hour shifts.
lOO-ton Cyanide "Plant Treating Concentrates. — This would include the superintendent and his assistant; tiie accountant and the chemist; the foremen for each shift at the mill and for each shift of the refinery; also the foremen in chai^ of repairs; in the mill the solution, filter and pachuca men; in the refinery the refinery men; finally, under the repair boss, his repairmen, repair helpers and the common labor (often called rouatabouts).
THE PURCHASIKG AITD SELLIITG DEPARTUENT This attends to the purchase and delivery to the works of ores, fuel, fluxes, and general supplies for repairs uid renewals. It may also attend to the sale of the products of the works.
The Purchase of Fuel and Fluxes. — In Utah coke may be quoted at $12.60 per ton, f.o.b. at the works, at Pittsbui^ t5. Attention should be paid to its contents in moisture, to its proportion of fines, and to its analyms, and especially to the amount and constitution of its ash. So much does it vary in this regard that the buyer should be well informed in regard to the various makes of coke offered for purchase, and not depend on the price alone.
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506 Organization And Operating
Western coal may be quoted at $5.50 per ton for run-of-mine. For certain work alack-coal ifi quite suited. The favorable qualities are low aah, and the production of a long flame in the reverberatory funutce, a quality not so important for the boiler.
LimeBtone. — The price may be given at $2 per ton at Utah works. For fluxing it should be low in silica. It should not be friable nor contain much fine.
Iron Ore. — Eight dollars per ton at the works on a bads of 47 per cent iron excesA. An allowance or charge of 20 cents up or down is made from this figure.
The Purchase of Siqqrlies. — These consist of iron and steel, castings, tools, pipe and fittings, oil and waste, brick, clay, qtiicklune, and chemicals. The list price of most of these are given in the catalogues of supply houses and discounts from the list are given. It is well to obtain competitive bids for furnishing these; for those thinf^ to be obtained on short notice, the buyer can, by arrangement, obtain the usual discounts. These supplies are kept in a storeroom, and should be issued by the supply department only on a written order from the foreman, or other responsible person who needs them. In this way it is known where they are to be distributed on the cost sheets. An account is kept of all supplies received and issued, so that, from it, can be learned how much and when to order such material, to maintain the stock. It is detrimental to the business to so nm out of supplies as to cause delays.
Much knowledge is required in tiw purchase of supplies. The nile is to buy when prices are low, or on a rising market, but only in small quanti-. ties on a falling market, and to obtun the best discounts, taking care, however, to avoid the purchase of inferior goods.
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Chapter L
Profits And Costs
Profits
Profits ttom the operation of a metallur^cal plant, whether a mines I^ant or a custom plant which has to buy ita ores, may be defined as the difference between the total coste and the returas on the metal product sales. How varied are these costs is well shown in their enumeration under bead of " Accounting," pf^^ 592. ProEts may be increased by full opera-
Costs And Profits
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When Runninq It The Specified
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Fia. 330,— Coetfl and Profita.
tion, by better extraction or recovery from the ore, by economy of treatment due to methods permittinii; a saving of labor, supplies, or fuel, and by faster running by which output is increased.
The above graphic table shows how profits increase as the plant s its full capacity; it also shows that, in this particular case, profits
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S98 Profits And Costs
oeasT at 25 per cent of capacity and again that the profits per ton increase as full capacity is approached.
Custom Smklteries
The profits ot a custom silver-lead smeltii^ works are obtwned t^subtractii^ from the money realized by the sale of metals recovered, the total ccets for treatment, freight, refining, interest charges and selling costs.
Milling Ores. — In milling the calculation lemains the same wheUier the ore is highly silicious or not.
The following figures represent the profits of a company owning a mine, the Robinson company, on the Rand, South Africa:
Ch^ reoovOTcd at tbe stompa 120.60
Cost <rf miniiiK $6.65
Coat of miBing 0.98
Net profito per too $17.70
C<Bur d' Alone District.— Out of 336,630 tons of ore mined at Uie Bunker Hill & SuUivan mine in 1906 there were shipped to the smelting works 86,640 tons of concentrate or one ton in 3.84. This averaged 45.8 per cent Pb and 18.78 oz. Ag per ton. The ore, as mined, assayed 13.32 per cent Pb and 5.89 oz. Ag per ton, the loss by concentration being estimated at 10.43 per cent Pb and 17.06 per cent Ag, or 11.96 per cent of the comlnned product. Taking the average prices at 4.6 cents per pound for the lead and 60 cents per ounce for the sUver, we find the costs and profits per ton aa below:
Assay value $15. 78
Freight, treatment 3 . 71
Smelt«r deductionH 3 .08
Average Profit »4.S7
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Index
Acceasoriei of the blaet-f unuu», 298
Aeooimting department, 592 Accoimte, coet, 604 — , distribution, 602 Add, BeaBemer process, 321
lined converter, 322, 323
operation, 322
— parting of silver-gold bullion, 278
— refntctoriee, 31, 32
— treatment of i ino-boz precipitate, 183 Actioa of machine in crushing, 60 Administration and general charges, 59i
— deportment, 684 Ac^meration of fine ores, 286 Agitotion of slime, 167, 160 .
— treatmoit, 161
Agitators, combined pneumatic and
mechanical, 160 — , data of, 187 — , mechanical, 15B — , pneumatic, 16S Air, etmibustion in, 76 Ajo process, 432
, letLching tanlcs for, 435
Akins daseifier, BO, 70
pigu
Alaska TVeadweU mill, 215, 216
Alkaline earths, action of in slags, 482
Allen cone classifier, 69, 70
Alloy steel, 347
All sliming cyanidation, 157
Alumina in ilagfi, 482
Aluminum dust, precipitation by, 261
Amalgam, safe, 241
Amalgamation and oonoentration mill, 246 of silver ores, 234, 244
— (patio process), 249 American filtering machine, 170
— ore hearth, 464 Ammonia leaching, 438 Anodes and cathodes, 451
— — — , tiftTiHlmg^ 463 — , sampling, 49 Anthracites, 12 Antimony gold ores, 146
— in silver-lead smelting, 483 Arbitration, 682
ArgaU roasting furnace, 97 Arwwic in Nlvo^lead smelting, 483 Arsenical gold ores, 145 Artificial fuels, 11 Association of employers, 683 Auguitin i»x>oe«s of silver milling, 2S0 Automatic or machine sampling, 42
Bag bouse, 488 BaUmiD, 66
, Hordinge, 67
in closed circuit, 66
, path of travd in, 66
, proportions and efficient, 66
— or tube-mill drive, 63 Bar screens or grissly, 66 Barrell chlorination, 137 , plant for, 141
Base bullion, costa of refining, 611 ~ — furnace, 601 ^ — refining, 498
sampling, 48
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eoo
BaK bullion, nHmming of, 608
— metal ores, 4
Bbbbh, action of in aUgs, 481 Bnsic fieaaemer proeem, 326
— copper converter lining, 402
— iron, 313
— open-hevUi charge calculation, 336, 33tl , chemistry of, 337
operation, 336
prooeaa, 334
— refractorieH, 37 Batbiry, st&mp, 125 — , toi-stamp, 127 Bedding lead ores, 407, 499
— ewes, 40
— system, Menit«r, 361 Beehive coke, 319
Bd^um linc-Bmdting furnace, 629 Belmont Milling Co. flow sheet, 266
miU, 266
refinery, 288
Belt conveyor, 660 Betta prooeaa, 511 Bessemer pig, 315
Black copper smelting, 359 Blanket or canvu tube, 114 Blake crusher, 61, 62 Blast, dry air, 305
— or pot roasting of oree, 110, 286
— roasting of galena, 110 pot, 111
Blaet-fumace, 72, 287, 289, 290, 291, 292 298, 306
— accessories, 367
— burdening, 310
— copper matting, 369
— heat-balance, 300
— , iron and acccesoriea, 298 — , — , operation of, 300, 302
— irregularities, 302
— operation, 371
— plant for copper ores, 355, 356, 369 fOT ii»n, 287, 288
— , showing temperaturea, 305 — , silver-lead, 471, 472. 473 — , — , products of, 487
Blaat, smdting and converting plant, 408;
— tap jacket, 474
— tuyere, 274
— St. raverbemtOTy, 386 Blende, rotating of, 510 Kistet copper refining, 442 Blower, positive blast, 368, 369 Blowing engine, 299
silver-lead blast-fumaoe, 476
tbe iron blast-sumacs, 301
Bono ash, 37
Both and molds, 8
Bms process of silver milling, 243
Bottoms, treatment of, 389
Box, xinc or extractor, 181
Braun disk grinds, 42, 44
Brickkiln, 33
— making, 38, 36
— mold, 35
— repressing machine, 36 firicke, 32
Briquetting flue dust, 489 -.-ores, 287 Bmmo-eyanide process, 140
O'Haia roasting furnace, 97
Brunton's quartering shovel, 42
— roasting furnace, 107, 108 Belt elevators, 567
Bins and pockets, 563
Bitumite, 12
Buchut concentrating table, 115
Buggy, charging, 551
Burdening the iron blast-fumace, 310
Business of metallurgy, 578
Butte-Duluth mill, 431
— process, 430
Butler's vacuum-leaf filter, 168
— filter frame, 169 By-product coke, 20, 21, 22 idant, 22
Caldnation, 88
Calculationa of charge in pyritic smdting.
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CalculationB of charge in Teverbera
ameltiog, 3B7 nlver4ead smelting, 484
CaJdecott diaphragm cooe, SB Callow flotation maohine, 117 Caloriraeter, Mahler bomb, 80 Catomnctiy, 86 Canvas table, 114 Capacity of roasting furnace, 100 Capital expenditure, 502 — involved, 591
Carbon brick, 38
— in pig iron, 316 Carbonate iron ores, 2S5
Carriage and teet, English cupelling furnace, 600 Cars, hopper, 664 — , industria], 663 — , aide-dump, 653, 664 — , transfer, 654 Casting copper, 8
— gold, 7
— iron and steel, 8
— lead, 8
— machine, endless mold, 414, 445 , Walker, 447
— silver, 7
— sine, 9 CastingB, steel, 348 Cast iron, 318
— — nioid, 318 Centrifi^ beh, 507 Chamotte, 36 ChatBcteristics of lead ores, 461
silver ores, 231
Chimneys or stacks, 78 Charooal, 16
— in ailver4ead smelting, 484 — , pig iron, 316
Charge calculation for the iron blast-furnace, 310
add open-hearth process, 333
basic operh4earth prooeas, 330
iron blast-furnace, 312
reverberatory smelting, 397, 398
— , reverberatory copper furnace, 303
— sooop, line smeltdng, 534
Chai^^, administrstion and general, 504
Charging buggy, 561
— lead ores, 567
— machine, copper, 446 , open: hearth, 333
Chemical reactions of the basic open-hearth process, 337 copper converter, 404
in the silver-lead blast-furnace, 479
ChNnistry of roasting, 89, 92
the cyanide process, 146, 264
Chilian mill, 60
Chloridiiing blast roasting, 249
— roaster, MacDougall, 428 , Wedge, 424
' — roasting, 88
Chlorinstion panel, 137, 138, 139
— mill, Goldfidd, 136
— of concentrates, 135
copper ores, 422
gold ores, 135
— or Plattner process, 133
— , precipitation plant for, 141 Chromite or chrome iron ore, 34 City Deep mill, 191 Clarifying solutions, 176 Classification and definition of ore, 3
— of iron ores, 283
metallurgical operations, 5
natural solid fuels, 13
pig iron, 314
Claasifier, 68
— , Caldeoott diaphragm cone, 60
— gold mill concentrates, 213
— of gold ores for milling, 122 Clayey gold ore, oyaniding, 189 Cleaning iron furnace gases, 295 Clean-up, 120
— of sine boxes, 182
— pan, 128, 120
Coal-4red reverberatory furnace, 301 Coals for roasting, 14
Co>J> in the United 6t«t«e, compoAition
of, 13 CtMkne grinding, 66
— or jKimuy cnHhing, SI, 285 Cobalt milling practice, 274 Cdka, 18
— Mh,18
— in ailver-lead smelting, 4S3
— oven, by-product, 20, 22
— plant, by-product, 22
— pusher and leveler, 22 Colorado lead ores, price of, 483, 496 Combined mechanical and pneumatic
agitators, 160 Combustion, 75
— in air, 76
the blastr-fumace, 76, 77
— of fuel, 81
— , imnciplee of, 75 — . temperature of, 79, 82 Comminution of oic, S2 Comparative agitator data, 187 Composite costs, fi7S Composition of copper matte, 372 Comprened-air locomotive, 6S2, 563 Concentrate, chlorication of, 136 — , sampling of, 46
— treatment at Goldfield Cons, mill,
218 Coitcentrating table, Burchart, 116
. Wilfley, 115
Concentration, 114, 26S
— in stamp milling, 131
— prior to cyaniding, 143 Concrete floare, unit costs, 573
— foundation, unit costs, 672 Condensers for linc smelting, 636 Conical mold, 8
Coning and quartering, 41 Consolidated I^n^aaghte mill, 193 Construction of plant, 550 — , unit costs of, 572 — , winter work, 570
Continuous counter-current decantstion, 166, 168
— grinding pan, 237
— handling of materials, 557
— thickener, Dorr, 163 Convenion of white metal, 405
Converter, acid-lined, 323
— and mixer building, 326, 341 — , copper, Smitb-PieMe, 401, 403 — , — (upright type), 401
— , electmlyticaHy opovted, 323
Converting copper matte, 400, 403, 405
— leady matte, 492 Conveying, 365 Conveyor, belt, 569 — , discharge, 560 — , endkee chain, 661 — , incline to level, 661 — , screw, 569
— tripper, 561
Copper and copper ore prices, 457 bearing gold ores, 145
— blast-fumace conditionB, 360, 371 I^ant, 356
smelting of ondiied ores, 355
slags, 367
machine, 414, 445, 447
— charging machine, 446
— converter, 400
— converting, losses in, 405
— electrolytic tank, 460
— extraction from its ores, 352
— furnace slap, 374
— in silver-lead smelting, 483
— leaching, Ajo process, 432
, ammonia proven, 438
, Butte-Duluth process, 430
, Henderson process, 423
, lAist process, 426
plant, 440
, Rio Tinto process, 418
— matte, 373
, composition of, 374
converter. 400, 401, 402, 408
, converting, 400
— matting blastfurnace, 363, 364, 36S
— operation, 371, 373
— ores and their treatment, 351, 417 , chtunctcnstice of, 351
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Copper ores, hydrometalluigy of, 354 , pyrito Bmelting, 354
— — , ameltiog for matte, 353 — , piopertia of, 353
— refining, 442, 440
, electrolytic, 449
furnace, 443
— reverberatory emehing, 387 Coats aCEounla, 594
— and profits, 597 — , Qompoeite, 575 — , general, 592 — , kinds of, 593
— in cine Bmelting, 538
— of sdminiBtration, 502 agitation, 187
concentration in nullinf;, 132
copper furnace plant, 414, 415
refinery, 465
operation, 4S5
' cyaniding on the RAnd, 155
-— ailvec-bearing concentrate, 277
dinolution by slime agitation, 186
filtration or decantation, 187
lead ores, 494
making coke, 23
Homeatake plant, 156, 198
operating copper works, 414, 416
plant and equipment, 669
steel, 348
Rand cyaniding plants. 156
rebuilding, 570
refining base-bullion, 511
roasting, 109
Bilver ores, 280
fllime plants, 186
treatment, Belmont-Tonopah mill,
, Bomeetake mill, 227
, Modderfontein mill. 227
Unit«d Eastern mill, 211
Cottrell treater, 400, 413 Counter-current agitation, 165 Cowper hot-blast stove, 297 Crane ladle, 406 — , track (and grab bucket), 566
Crane, traveling, 556
Crowe vacuum proceaa, 177
Ciusbei, Blake, 51, 53
Crushing, action of machines in, 60
— and screening, dry, 64
— , grinding, screciiung and daasif ying, 60
— rolls, 58, 50 Cupelling furnace, 506
— rich lead, 507
— nlver precipitate, 262
Current flow, electromagnetic. Walker multiple ByBt«m, 451
, testing, 464
CiMtom smelteries, 598 Cupola furnace, 73, 77
, combustion in, 77
Cyanide process, 133 , chemistry of, 146
— solution, strength (rf, 146 Cyaniding concentrator, 221 — , double treatment in, 113
— free milling ores, 189
— mixed silver ores, 268 — , outline of process, 143 — , sand leaching in, 151
— , silver-bearing concentrates, 277
— silver ores, principles of, 256, 264 — , systems of, 150, 223 Cylinder drier, 100
Daily wages and premium, 500 Decantation in cyaniding, 167 — , continuous counter-current, 165 — , costs of, 187 ■ — , intermittent, 168
Definition of metallurgical thermochemistry, 84
straight or simple ores, 3
Definitions and cUssificalioD of ores, 3 Dehne fifier prem, 171 Deister concentrating table, 116 Department, administration, 584 — operating, 586
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D^MTtmeDt, puTchaBJDg and seUinf, S9S
Depreciation, 502
Desilveriiiog base-bullion, 276
DesulpburiiinK piwxm at Nipiaeiiii, 276
Denncing furnace, 605
Dip B&mplea. 46
Direct prooeee of reverfaeratory smdtiag,
390 Diaks of Symons crusber, 62 Diapoeal of pig iron at iniii blaflt-fumaoe,
304 — ~ — slag at copper furnace, 384
— iron blast-fumaoe, 303
Distillation of linc, 640
— of line ores, 528 Dietnbution account, 592
— of wealth, S79 DokKDito. 37, 482 Dorr Hgitator, 160
— bowl clanifier, 69, 71
— daasifier, 56
— continuous ttuckener, 163, 164
, showing dine RDneB, 164
Double atrand elevator, 568
— treatment in cyaniding, 158 Drceaing the linc boxes, 184
— the plates of the stamp battery, 128 Dry air blast, 306
— crushing and ecreening, 64 flo„ sheet, 66
— silver milling (Reese River prooena), 249 Dryer, cyUnder, 100
Drying and refining the gold precipitate,
silver precipitate, 261
Duties of foreman, 586
superintendent, 585
Duplex and electric furnace plant, 3^1,
342 — ' proceM of steel tnaldng, 340 Dwight-IJoyd sinter roasting machine, 110, 112
ie situation, 576
in the United States as rdated t
metals, 580 Economics of engineering, 679 Edwards roasting furnace, 97, 100
Efficiency of men, 587 •'>
Electric fumaoe and control panel, 340
building, 344
for steel making, 342, 344
smelting of sHver precipitate, 262
showing lining and bottom connection
, 344
— lighting, unit costs of, 573
— locomotive, 551, 562
— steel making, 342
— tranaformer and substation equipment,
— trolley removing slag. 384 Electrically operated converter, 323 Electrolyte, circulation of, 454
— copper, 451
Eleotrolytk) cop^ier refinery, 44Q refiiung, 449
— putting of silver and gold buUion, 279
— refining of lead, 511
— tank, 450
Eleetrostatic Cottrell treat«r, 409, 410
— recovery of smelter dust, 410 Elements, heat of formation of, 87 Elevators, 566
— , belt. 557 — , feeding, 558
— , single and double strand, 558 Elimination of unpuriticB (copper converting), 405 Employers, association of, 583 Endleai-chain conveyor, 561, 562
mold casting machine, 445
Engineering, economics of, 579 En^ish cupelling furnace, SOS Equipment, obsolete, 593
— of plants, 551 KxcavatoiB, 555 Expenditure of capital, 592 Extraction of copper as chloride, 422
smelting, 353
from ores, 352, 353, 418
Extraction of gold with solvents, 133, 134 silver from ores, 231
Faber du Faur retortit^ fumaoe, 6( tilting fumaoe, 363, M8
.''eed, removal of wood and tramp iron
from, 565 Feedere, ore, 583
— , traveling, 664 Filt«T, Amoricsn continuouB suction, 17D,
— frame. Butters, 169
— , Oliver, continuous suction, 170
— prwa, Dehne, 171, 172, 173
, Sweetland, 173
— atime treatment, 157 Filters, general remarks on, 175 Filtration, 166
Financial crises in the Vnit«d StAtcs, 683
Finishing the sample, 45
Fireclay, firebrick and tile, 34
Flame temperature, 80
Flotation, 116
— machine. Callow, 116 -, Janney, 116
, Minerala Separation, 116
Flowsheet of AlaBka-Treadwell mill, 216
Hollinger mill, 216
Waihi Grand Junction mill, 271
Rue dust, briquetting, 480
, treatment of, 489
Fore-hearth or settler, portable, 366, 367
Foreman, duties of, 586
Foundry pig, 315
Fractional sdection, 41
Free milling ores, milling of, 189
Ftenier pump, 566
installation, 668
FVue Vanner, 116
Fuels, 11
~ and fluxes, purchase of, 695
— m silver-lead smelting, 483
the open-health furnace, 332
Fuels, natural, 11
Furnace and stoves iron, 208
— , hand reverberatory, 74, 9S
— shaft, 72
wind, 72, 77
Galena ares, 461
Ganister, refractory, 32
Gases, metals and physical constants, 6 — , specific, heat of, 80 Gasoline locomotive, 562, 653 Gate for ore bin, 561
General arrangement of gold dtomp mill, 130
— economic situation, 579
— remarks on filters, 176 Genesis of natural fuels, table of, tl Gjters calcining kiln, 286
Gold, amalgam retorting, 129
— and silver, valuation of, 122
— baiB or ingots, somiding, 48
— casting, 7
— mill ooncentrstcs, treatment of, 213 ■, the City De^, 191
— milling practice, 189
— ores, amalgamation <rf, 123
and classification for milling, 121
, arsenical, 145
, copper-bearing, 146
, graphite, 146
, hydrometalluigy of, 133
, ocourreoce of, 121, 133
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fl06
Gold orea, prio«a, 227
, pyritic, 145
, ailioioua, 14fi
, amelting, 228
, taktoae or day«y, 146
, tdluridn, 145
— , physical iHopertiM of, 121
— telluridai, 121
Golden Cycle flow sheet, 201
mill, 201
Goldfidd chlannatioa mill, 136
— CoDBolidatod mill, oonoentnte trait-ment
at, 218 Qmb-buoket and track enoe, SAB Grabe and ezcavaton, 555 Grading oraa, 5
— pig iron, 316 Graphite, 13, 16, 34
bearing gold orea, 223
Great Falls oc^per converter, 401 Grinding and dasnfying on t^ Rand, 192 — , ooane, 65
Gyiatory crusher, 52, 53, 54
H
Hammer feeder, 465
Hand rererberatory fumaoe, 74
— sampling, 41
Handling of materials, oontinuotH, 567 Haidinge conical bfdl mill, 57 Heap roBstiiig, 92
trf matt«, 94
Hearth and boah, iron UastJUinace, 202 Heat balance of the blast^umaoe, 309
— evolved in roasting, 87
Heate of formation of the elements, 86
, table of, 86, 87
Hegeler roasting furnace, 97, 621, 628
Hematite, 283
Hendnwia process, 423
Hwdryx agitator, 160
Heyl & PateraoD pig-casting-machine, 304
High-grade Nipissing mill, 243
Hoists, industrial, 553, 564
Holbeck powdered coal syBt«m, 30
HoUinger miU, 206
flow sheet, 206
Homestake mill slime plant, 197 trafttment costs, 198
— plant (cyanide), owt of, 156, 198 Htq^wr can, 664
Horisontal coiqwr converter, 400
, Pieroe-Smith type, 403
Hot blast stove, 295, 297
— mfltal mixer, 321, 322
— pnm, 604
Busies gw producer [dant, 25 Hydrometallurgy of copper ores, 354, 417
gold orea, 133
sQvw orea, 260
HypoeulphiU lixiviating of ailver ores (Von Patera prooeaa), 254
InqMct somen, 66, 67 lu^Hirities, elimination of in oopper cooverting, 409
— in metals, 6 tnadequato equipment, SS3 Incline conveyor, 561 Indicating pyrometer, 83 InduHtrial cars and haista, 553
Influences of elements on pig iron, 316 Ingot copper, section of bar, 49
— mold, 324
Inside men, morale of, 688 Intermediate Gruahing, 56 Intermittent deoantation, 168
— hm^^ing of roaterials, 561 International Smelting Co., reverberatorv
, roaster plant, 411, 412
Iron and steel, 281
Iron blast^umace, 287, 280, 290, 291, 282,
charge sheet, 312
, cbamical reactions of, 306, 307
, detailed section, 291, 305
plant, 287, 288, 289, 205
, showing t«nperaturee, 306
with automatic ol
Iron ore, action <d in sbtgs, 481
and itfl smeltaiig, 2S3
, olasoflcation and oocunence, 283
, pricea of, 347
roasting, 2Sfi
Irregulajities of blast-furnace opontion.
Jacluon centrifugal pump, 567 Janney flotation machine, 117 Jones sampler, 42
Kalgooilie district, 203
KeUy filt«r prm, 171
Kennicott plant for Ammonia leaching, 440
Kiln KMating, 86, 286
Labor situation, 680 — , union and non-union, 680 — , unit Msta for, 604 Udlecan, 367
— crane for copper, 406
— , worm-geared, bottom-tapped, 324 I«irtr prooesB, 426
mm, 426, 427
Large copper matting blast-fumace, 36B
, Butte-Duluth proccHB, 430
ores, 417
^ plant, Henderson process, 423 . I*irt process. 426, 426, 427
— the aanda in cyanidii% 1S3
— vats in cyaniding, 151
Lead, blast-furaaM, 471, 472, 473, 476 — , — , charge, 484 — , casting, 6
— copper matte, 491 , treatment of, 491
— ores, carbonates, 462
, classes of, 461
, costs of smdting, 494
, ondiied, 482
, penalties on, 496
, prices, 496
, piopertiee of, 461
Lead ores, receiving, sampling, and bedding, 467
— refinery, 499 Liberty Bdl mill, 199, 200 Umestone for sla^^ 482 lining of copper converter, 402 Location of i^anta, 547 Locomotivea, induBtrial, 651 Long-hearth revarberatory roaster, 96, 96 Loomis-PettiboDe gas apparatus, 27 Lasses in oopper oonverting, 406
sine smdting, 637
M
McArthui^Forreet proceaa, 133 MacDougall roastii^ furnace, 101, 102-
104 Machine or automatic sampling, 42 Maclmery, costs of creation, 673
— prices, 671 Magneaite, 37 Magnetite, 284
Magnetised pulley for removal fA tramp
iron, 566 Mahler bomb calorimeter, SO Main systems of filtering, 168 Maintenance and repaiia, 693 Making of atAel, 320 MaUeable iron pig, 315 Manganese, action of in sla^, 482
— in pig iron, 316 Manufacture of steel, 320
wrought iron by puddling, 318
Market lead, molding, 606 Martin sampling machine, 47 Materials in intermittent handlteg, 661 Matte (copper), 373 — , he^ roasting of, 94 — , leady copper, 491
— smelting charge, 378, 381
of copper ores, 364, 360, 371
Mechanical agitatora„ 150
— open-hearth chargmg, 333
— roasting furnaces, 97
Melting and refining I^dce oopper, 444 Mo), caie of, 588
H«D, effideocy of, 587 Merrill filtar pnm, 174
frame, 174
iiuUll&tiaa, 173
plate, 174
— precipitation appAratus, 178 prooew, 178
Mercury fed to the stomp battery, 126 Merton roMting fumace, 07, fi26, 626 Meentm bedding system, 361 MetaUurgical fumaoea, 72
— opemtioiui, daanfioatione of, S
— thermo-chemistry, 84
, claasification tA, 84
, imitfi of measurement, 84
Metals and ptaea, jAymai oonatanta for, 6
Methods of treatment, 4
Mexican amalgamation process, 249
Mill amalgamation and concentration, 246
— iron, 314
— , Nipissing Co. 'a low-grade, 274
high-grade, 243
— samples, 46
— sites, 54B — , stamp, 65
Milling, dry silver, 240
— gold ores in stdutions, 133
— ores, 5B8
— practice at Cobalt, 274 MiUs, typical sflrer, 266
Minerals Separation flotation machine,
lie
Mixed gas jnixlucer, 24
— ores, definititHi of, 3 Mixer, hot metal, 321, 322, 341 Mixing pan, 324
Modes of paymmt, 680 Moisture samples, 40 Mold, cast-iron, 8 Molding and casting metals, 7 Multiple-hearth furuare, 101
Native copper, 351
— gold, 121
Natural draft furnace, 76
— fuels, 11
, genesis of, 11
— B»s. 15
— soUd fuels, 12
, classification of, 12
Ne^ect of equipment, 503 Neutral refractoriee, 31 Newnam ore hearth, 465 Nipining Co.'s low-grade mill, 274 mill, flow-flheet of, 275
wet desulphuriting process, 276
— mill, hi^-grade, 243 Noduliiing iron oree, 286
O Obeolete equipment, 593
of gold ores, llfi
Oil flotation, 116
Oliver continuous filter, 170
Open and closed-top blast-furnace, 475
Open-hearth furnace, 327, 328, 329, 330
, charging, 333
fuels, 332
, reversing valves tor, 331
, steel-making in, 326
, tilting, 332
, with water-<fooled device, 329
, chemistry of, 337
. recarburiiation in, 337
Operating department, 585 , cyanide plant, 595
, gold stamp in ill, 595
, silver lead or copper smelting, S94
Operation and organiiatioa, 684
— of basic converter, 403
copper blaat furnace, 371
plant, 587
diver lead blast-fumaoe, 476
the acid-Jined converter, 322
iron blaat-f umace, 300
rinc smelting furnace, 434
Ore bins and pockets, 563
— feeders, 563
— gate, 564
— heart, 464
— storage and supply, 563
— treatment, 593 Ores and metals, 3 — , base metal, 4 — , copper, 351
Organization and operating, 684
Oroyo-Brownhill miil, 204
Outline ot process of cyaniding, 143
Overstrom concentrating table, 116
Oxides and carbonates of oopper, 362
OxidiiiDg roasting, S8
Oxland roasting furnace, 97
P
Fachuca tank, 15S
Fan, dean-up, 129
Farkes process, 503
Parting silver and gold bullion, 278, 279
Path of travel of ore particles, 66
Patio process of silver amalgamation, 249
Payment, modes of, 589
Penalties on lead ores, 496
Petroleum or fuel oil, 16
Phosphorus in pig iron, 310
Physical properties of gold, 121
Piece work, 589
Pig casting machine, 304
Kg iron, 314
, classification of, 314
, costs of production, 348
Pig iron disposal, 304
grading, 315
, influence of elements on, 316
prioea, 347
sampling, 49
, smelting for, 287
Pittsburg pig iron, 315
Plan of furnaces and stoves, iron, 298
United Eastern miil, 210
Plant for dujriex and electric furnace, 341 dnc smelting, 533
— operation, 587
Plants, construction of, 660
— , equipment and their costs, 546
— , metallurgical, costa of, 571
Plate amalgamation of good ores, 123
Ptattner process, 133
Plumbago, 34
— crucibles, 8 Pneumatic agitators, 168 PorUble fore-hearth or settler, 366 Positive pressure blower, 368, 369 Pouring ladle, 324
— slag, 373
Power needed in crushing, 60
Practice in roll crushing, 50
Precipitate, blast-furnace smelting of, 188
Precipitation by aluminum dust, 261
— of gold from solutions, 177
sUver from cyanide solutioos, 257
Preliminary construction work, 670 Premium and daily wages, 590 Preparation of ores, 39 Pressure filtration, 108, 171 Prices of iron ores, 347
— — lead-silver ores, 464, 496
pig iron, 347
raw materials, 576
silver and silver ores, 280
steel, 347, 348
sine ores, 539
Principles of combustion, 76
samiding, 39, 46
sUver, 2fi0
Priocipleg dating to the rafiniog of
met&ls, 6 Produoer gas, 23
, coete, 28
, muted, 24
, idant, 26
, aiiaple, 23
PHtduetion and eaeta, 601
prioea of lead-silver orea, 4M
Products of Bilver4ead blast furnace, 487 Profit'^haring, 690 IVofita and costa, fiOfi
— of treatment of mlver^ead orei, 496 PrDpertKB of copper, 363
PnqmrtionB and efficnency of ball mill,
— proMH,31S
, reactioQB in, 319
Pug mill, 490
Pulp Bunpler, 47, 48 Pulveriied coal, 29
distributing syatem, SO
fired revnb«nttory furnace, 392
in eilvBr4ead Bmalting, 484
Purehase of fuel and fluxes, 596
Purehasing and adding deparbnent, 695
Pyrite gold ores, 145
— smelting, caloulatkin of chaige, 381
of eloper ores, 354, 377, 379, 383
, section of furnace, 379
Pyrometer, indicating, 83
Quartering an ore in samp
— and coning, 41
— shovel, Brunton's, 42
Rand oyaniding costs, 156
— practice, steel tanks for, 194 Raw material prices, 576 Raymond roller-mill, 29
Reactions in copper reverberatory smelting, 397 puddling process, 319
Reactiona in sinter roasting, 113
Realisation charge, 694
Rebuilding, cost of, 670
Recaiburiaation in opm hearths, 337
Receiving ores, 40
— , sampling uid bedding kwi ana, 467
a62
Reoovery of BtatitBt dusts, 410
sulphuric acid, 527
Reduction charges, 693
Reese River dry process of milting, 249
Befinoy, lead, 499, 613
— , Bants Gertrudis, '280
Refining f umsce for copper, 443
— in a cupelling furnace, 184
— lead and base bullion, 498
— of ooppw, 442
metals, prinraides relating to, 5
— sUvet precipitate, 259, 261
— with potassium bichromate, 184
— sine, 640, 642 Rdraotories, 31, 38 — , add, 31, 32
Refractory materials, inoperties of, 131
Regular operation of the iron blastrJuniMe,
302 Rents, taxes and royalties, 6M Repair and maintenance costs, 593
— force, 586
Reports, 694
Repressing machine, bride, 36
Requirements of capital, 691
Research work, 686
Retwt and furnace for silver mill, 2^
— manufacture, 636
Retorting furnace, Faber du Faur, 508
— gold amalgam, 129 Retorts for sine smriting, 635 Reverberatory coppw smdting, methods
of, 386, 387, 390, 397 , Wdah procM, 387, 3S8, 389
— fumaoe, 73, 74, 327, 387 Reverberatory furnace, opm-bearth, 327
— roaster, long hwrth, 96, 96
— smelting and converting [dant, 412, 413 , direct {koccsb, 300
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Reveiberatoiy ameltine funuwe, 387
Roasting inm ores, 286
, charge, cKlciUationB, 3B8
,owa-fiml,391
— of nmtte, 92
, method of charging, 3B2
— ores in pulverised oonditioD, 94
, oa-fired, 394, 3SS
— in. bUst-turnace Bmelting. 388
Reversing vbIvgb, open-hearth furnaoe,
— line ores, 619
, chemistry of, 619
Robbins-MeeaitcT reclaiming machine,
— screen or trommel, 68
Roller mill, Raymond, 29
Rich lead, 507
Rolls, crushing, 68, 69
, cupening, 607
Roofing, unit poets of, 674
, tw»tment of, 607
Ropp roasting furnace, 97
Bidge Touting funiace, 626, 626
Rotary feeder, 464
Rio TiDto prooew, 418
■
Rules of works, £87
S
Run-of-mme ore, bim of, 60
Samide finishiog, 45
RusBell proccHB for silver ores, 264
— grinding miU, &
Roaster plant, 106
— miU, 46
Roasting, 88
Sampler, Jones, 42
- Hende, 619
-.pulp, 47
pulse,46
— ftmiace, AripH, 97
— 0OM»er ingots and anodes, 49
, blende-, 521, 622
, Brown horsEsboe, 98
, Brown-O'Hara, 97
— iron ores, 46
, Bninton, 107, 108
— lead ores, 467
, cylinder, 98
— machine, Martin, 47
, Edwards. 97, 100
— metids, 48
— miU, 43, 44
-oiw,40,286
, Merton, 97, 524, 526
- pig iron, 49
, Oxland, 97
, Pierce turret, 98
,Ropp,97
. Wedge, 103, 106, 106
, Wethey, 97
— pUnt, Homestake, 147
, White-Howefl, 97, 99
— pump, Frenier, 568
— fumaoea, capacity of, 109
Santa Gertrudis reenery, 280
— in hJhis, 88. 285
the Welsh prooea, 889
Screen impact, 06
Dgnz..byG00glC
— tetiflion, 67 Senening, 66
Screw coQveyor, 669, 660 Scrubben for stove vid boiled, 206 Section of bar or ingot copper, 49 SeUinn depMlment, AB6 Semet-Solvfty by-product ooke oven, 21 Septra tioQ of slime from wdutions, 166 Settler, ei^Uoot, 240
— or far»4ieartli, portable, 366
Shaft funi&ce, 12
Shafting, pulleys, and belting, unit coat
for, 674 Shaking screen and feeder, 664 Shannon Copper Co. proccn, 421 Sharing profits, 600 Side-discharge car, 664
— -dump car, 663 Silica brick, 32 SilicatM of copper, 362 Silicious gold ores, 14fi Silicon in pig iron, 316
Silver bars or ingots, sampling, 48
— casting, 7
— ,' distribution of in a bar of base bullion, 40
— extraction from ores, 232
ores, prices of, 280
lead bag-house, 488
bUat-fumace, 471, 472, 473, 477
, chemical reacticns in, 470
smelting, 467
and refinery, 613
plants, location of, 648
, pulveriiod coal for, 484
orea, profits of treatmNit, 496
works, 468, 471
— milling, Boss process, 242 by AugUBtin procees, 260
^ hydtometallurgical prDceeaes, 260
Ziervogel process, 261
. dry, 249
, wet, 2S6
— milk, retort and furnace for, 243 , typical, 266
with tank settling. 236
— OHB and treatment, 231, 232
Silver ores, amalgwnation and oonoentntion, 244
, characteristkB of, 231
, ehloridising roasting of, 246
, bypoaulphita process, 254
— precipitate, drying and feSiung, 261,
263 •"-^ — I melting in tilting fumaoe, 262
, smelting of, 262
— , precipitation of, from solulicMie, 257
— refining furnace, Faber du Vkut, 263
, Monarch-Rockwell, 263
, Steel-Harvey, 263
Simple ores, definition of, 3 Singl»«traud elevator, 568 Sintw building, 470
— charge, 470
— roasting, 88, 110
machine, 110
, reactions, 113
Sintering ores, 470
Situation, general economic, 570
SisM of screens, 51
Sising teats on sand residue, 1S4 Skimmer, base bullion, 605 Skip car, 655 Slag pot, side-dumping, 478
, two-wheeled, 367
— , actions of bases in, 481 — , ladle and locomotive at iron fumam, 303
— or cinder diaposal at iron furnace, 303 ", pouring (copper blast-fumaoe plant),
- cake, partly removed from filter, 170
- plants, costs, Homestake, 198 , — of, 186, IBS
-, separation from solution, 106
- treatment by filter, 157 Sliming, all, 167 Smelteries, custom, 698 Smdting by the Wdeh process, 388
~ coppca' ores, 353
- on the ore hearth, 463
- of copper slag, 357
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Smdting of gold ores, 228
lead ore, 463
ail w-le«d OTBB, 467
— plant for copper oree, 355, 366
— silver precipitate, 282
— to black copper at Union Miniere du
Haul Katanga, 367
— VB. cyaniding gold orei, 226
— sine ores, 519, 528 Smitb-Pierce copper converts, 401, 403 Smoke damage, 649
Solution, coats of by slime agitBtion, 186 Solutions, clarifying of, 17S — , precipitation -of gold from, 177 Sorting or pickling ore, 285 Specific heat of gases, 80 Split-shovel sampling, 41 Stage grinding, SO Stamp battery, 125 Stamp mill, 56, 124
' using amalgamation and concentration, 245 with plat« amal^mation, 124
— milling, concentration in, 131 Stamping, 265
Starting sheete, 462
— the copper blast-furnace, 871 Steam locdmotive, 561, 662 Steel, alloy, 347
— leaching vat, 152, 192
— making, 320
1 by t^ acid BessemK' process, 321
— — in the open-bearth furnace, 326
— prices, 347, 348
— rails, 345
— , structural, 346 — , tool, 346
— , wrou^t iron and, 318, 345 — , varietiea of, 346 Storing area, 40 Stove, hot-blast, 296, 296, 298 — , scrubbers for, 296 Straight ores, definition of, 3 Strength of cyanide solutions, 147 Structural steel, unit costs for, 574 Sulphide copper ores, blast-furnace smelting of, 360, 377 Sulphides of copper, 351
Sulphitixing toasting, E
npigu
Sulphuric acid leaching, 420
, Ajo process, 432
, Butte-Duluth proccw, 430
recovery, 527
Superintendent, duties of, 686 Supplies, purchase of, 506 Sweetland filter press, 172 Symons disk crusher, 61 Systems of cyanidation, ISO
Tailing, samplers, 46
Talcose grid orts, 146
Tank, Brown, 68
Tap-jacket, blast furnace, 474
Taxes, 694
Telluride gold ores, 146
— ores, treatment of, 109
Telpher mono-rail transport, 656, 656 Temperature of combustion, 79, 82
flame, 80
Temporary or distribution accounts, 692
Tension screen, 67
Ten-stamp battery, 127
Test for English cupelling fumaee, 609
Testing or leeearoh, 686
Thsmo-chemistry, metallurgical, 84
Thickening, 167
Thimble for bag house, 486
Three-throw plungm* pump, 66?
Tae,34
TUting fumaee, 8, 414
for copper, 414
, meJting in, 282
, Monarch-Rockwell, 263
, open-hearth, 332
, Steel-Harvey, 263
Tom Reed mill, 207
Tonnage in mills, calculations for, 223
Tool sleet, 345
Tramp iron, removal of, 466
Transformer and substation equipment,
344 Traveling crane, 556 Traveling feeder, 464 Traversing bridge, iron smdting, 287 Treatment by agitation, 161
— coetfl of load ores, 496
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TTMbnent, methods of, 4
— of «™alg«tTi, 241
battome, Webh prooen, 3B9
oopper ores, 351
Sue duBt, 489
gcdd-mill ooDcentratcB, 213
rich iMd, 607
precipitate, 269
tailing for acid or anuncMua leaching, 223 telluridc opea, 1B9
— or reductioti charges, 5fl3 Trent Mutator, 180 Triple roasting, 113
"niidex high-preeeure pump, 668 Tripper for conveyor, 660, S61 Tronunel op rerolving screen, 68 Tube mill, fl4
circuit, AlaslcA'Treadwell mill, 216
Unen, 63
— milling, 265
— or bfdl-mill drive, 63 Tuyere, blaaWumacc, 474 Two-wheeled tdag pot, 367 Typical gold mills, practice, 189
— silver millfl, 266
U UndereetimateH in building, 670 Union and non-union labor, S60
— Miniere du Haul Katanga, emelting at,
ase
Unit ooDBtruction costa, 672
— costs for concrete foundationH, 672
— erection of machinery, 573
electric lighting, 573
excavation, 673
labor, 574
masonry, 674
painting, 574
roofing, 574
shafting, pulleye, and belting,
ventilating windows, doors and
floors, 574 United Eastern mill. 570 United Eastern Mill, costs of, 211
flow-«heet, 211
Units of
istry, 84 Upright coppo' Utah lead ores, prioea of, 406
Vacuum Bltration, 168 leaf filter, Butten, II
m77 Valuation, 502 — of gold and silver, 122 Variations of poets of bnabnent of mlverleadorea
, 497 Various treatments and calculations, 223 Vats for cyaniding, description of, 1^
104 —, wUxA teaching, 164, 104. — , — , for double treatment of cyaniding,
154 — , wooden leaching, 152 Ventilators, windows and doon, unit coats
of, 674 Vesin sampler, 43 Victor plant, Portland Gold Mining Co.,
202 Victorious gold mill, 190 Von PatiOn, process, 264
, flow sheet, 271
Walker casting madiine, 447
— multiple system of current flow. 451 Wall-type indicating pynMoeter, S3 Wasp No. 2 mm, 189
Waste wood and tmnp iron removal, 465 Water jackets for copper blaBt^utnaee,
366 Wealth, distribution of, 679 Wedge blende roasting fumaoe, 522
— chloriditing furnace, 424
— roasting furnace, 622
Webb process of reverbastory amdting,
387 Wet dsBulpfaurising proven, 276 Wethey roasting furnace, 97 Wet silver mill, 236
milling with tank settling, 236
White briquetting press, 400 Whit»-HoweU roasting fumaoe, 07 -~ metal, conversion of ooppa, 406
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Wind funuoe, 72, 77
Winter work in oonstructioii, 670
Wood, 14
WotUng costs of lead orea, 465
Workmen, 686
Works, rules of, 687
Wonn or ac
Zinc boxes, dean-up, 182
, — and treatment of precipitate, 268
— coating, 10
— distilling, 540 — , grades of, 540
— in slags. 482
— or««,6l7
Zinc ores, reductioti of, 610
— or extractor box, 161, 182
— prices of, 639
— refill, 540
— smdting fumam, opa«tion of, 634 funtaMS, 629
, Belgium type, 629
, regenerative type, 632
, costs of, 638
, kMMB in, 637
, prices of, 639
retorts, 634, 636
Zones of the blast-fumaoe, 300
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