Cyanidation and Concentration of Gold and Silver Ores-John v.N. Dorr

Cyanidation and Concentration of Gold and Silver Ores

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Lundberg, Dorr, and Wilson mill, Terry, South Dakota, on Fan Tail Gulch-the scene of many new developments in cyanide technology. .

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Cyanidation and Concentration of Gold and Silver Ores BY

JOHN

v. N.

DORRIJ E.M., D.Se.

Member and Douglas Medalist ojthe Amm·. Insi. of Min. and Met. EngTs., Min. and Met. Soc. of Amm'., Inst. of Min. and Met., Can. Insi. of Min. and Met., Han. Member of Chem., Met. and Min, Soc. of S. AfT., Past Pres. Amel·. Inst, Chern, Erujrs., Member and Medalist of Soc. oj Chern: Ind., PeTkin Medalist. AND

FRANCIS L. BOSQUI Member of the ./imeT. Inst. of Min. and Met. Engrs., Chem., Met. and Min. Soc. of S. Aji·., Reg. Prof. Erujr., State of Conn., U.S.A.

III

II

SECOND EDITION

McGRAW-HILL BOOK COMPANY, INC. NEW YORK

TORONTO

1950

LONDON

CYANIDATION AND CONCENTRATION OF GOLD AND SILVER ORES Copyright, 1936, 1950, by the McGraw-HilI Book Company, Inc. Printed in the United States of America. All rights reserved. This book, or parts thereof, may not be reproduced in any form without permission of the publishers.

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I Foreioord When the senior author of this volume published his 'Cyanidation and Concentration of Gold and Silver Ores in 1936, no book literature on the subject had appeared since 1920. This barren gap of sixteen years was in marked contrast to the first two decades of the century, which were the prolific years for books on the cyanide process. In that period authoritative volumes on the subject averaged better than one each year, including two in the German language and one each in French and Spanish. The record reflects both the rapid growth and the ultimate establishment of cyanidation as the principal method of gold recovery throughout the world. A fortunate combination of circumstances lent timeliness and authority to Dorr's first book. Beginning in 1899 he was among the pioneer users of the cyanide process in the United States, gaining experience successively as chemist and operator, consulting engineer and plant designer. His mechancial inventions for classifying sand and slime, and continuously thickening the latter, were revolutionary and had a notable impact on metallurgical practice. Finally, his development of an engineering organization with world-wide branches and affiliates provided exceptional opportunity to keep abreast of practice and progress. It was out of this background and with these friendly cooperative facilities that he produced a work that found ready acceptance wherever cyanidation was practiced. It would be an oversight, however, to limit Dorr's reputation to his advancement of the cyanide process. His "cyanide machinery" embodied basic principles in classification and sedimentation that led to their wide adoption in scores of industrial processes. His inventions not only increased technical efficiency, but, in their application to municipal and industrial sanitation, immeasurably benefited social progress and human welfare. For his achievements' he has been honored with the John Scott Medal of the Franklin Institute, the James Douglas Medal of the American Institute of Mining & Metallurgical Engineers, the Chemical Industry Medal of the Society of Chemical Industry, the Perkin Medal of five joint American chemical societies, and the Modern Pioneers Award of the National Association of Manufacturers. Thirteen years have now elapsed since Dorr's book was published, and a critical revision was necessary before a new edition could be printed. This undertaking was entrusted to Francis L. Bosqui whose name now appears as junior author. Revision proved to be no light task and almost as time-consuming as preparation of the original volume. The result is v

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FOREWORD

practically a new book, an up-to-date manual of world-wide practice, particularly useful for consultation by operators and engineers. By a happy coincidence the junior author carries on the tradition of his illustrious father, Francis L. Bosqui Sr., "whose name and reputation "were widely known to an earlier generation of metallurgists. In 1894 the elder Bosqui abandoned the practice of medicine for which he had been educated, and cast his lot "with the Standard Consolidated Mining Co., Bodie, California, later becoming mill superintendent. There he started a career that led him successively to Colorado, Nevada, and South Africa as mill designer, operator, and consulting metallurgist. In his time he introduced the latest innovations in milling practice and equipment at the Liberty Bell, Smuggler Union, and Camp Bird in Colorado; the Combination Mines, and Goldfield Consolidated in Nevada; ModderfonteinB, and New Modderfontein in South Africa. While still at Bodie at 1899 he published Practical Notes on the Cyanide Process, one of the earliest books on the subject. With this rich heritage the younger Bosqui has followed in the footsteps of his distinguished father, both as metallurgist and author. He first worked with Mr. U. C. Tainton in the development of an electrolytic leadsilver process for the Bunker Hill & Sullivan Mining & Concentrating Co., Kellogg, Idaho. Thereafter he spent eleven years in South Africa, except for a brief interim in the United States, first on the metallurgy of Transvaal platinum ores, and later in research on Northern Rhodesian copper ores. With his father, who was consulting metallurgist to a group of South African gold mining companies, he made extensive investigation in the United States into methods of platinum concentration and extraction. As research engineer for the Rhokana Corporation at the Nkana mine he invented an electrochemical process for the production of high-grade cobalt and ferrocobalt from electric furnace alloy. He also developed the selective flotation of copper and cobalt sulphides. Other matters engaging his attention during this period included the elimination of bismuth from copper, production of oxygen-free copper in high-frequency induction furnaces, and the treatment of electrolytic refinery slimes. Returning to the United States in 1936, Mr. Bosqui joined The Dorr Company, and is now chief metallurgical engineer at the company's Westport Mill and Labora tories, Westport, Connecticut. Readers of this volume will be debtors, not only to the authors, but also to the fact that a policy of secrecy no longer dominates the profession of metallurgy. A book of this kind could not be written without the generous cooperation of the technologists who have brought the metallurgy of gold and silver ores to its present high state of efficiency and economy. Freely sharing their knowledge and experience for the benefit of all, they form an international fraternity of good will whose example might well be emulated

FOREWORD

vii

by social and political groups. But it has not always been so. Only a short fifty years ago the elder Bosqui wrote in the preface to his book: " .... the methods of operating, for which each operator claims a certain amount of originality, are, as a rule, secrets jealouslyjguarded, and in consequence we have a rather meager literature on the subject." . It is now generally recognized that free exchange of technical information and experience has been a weighty factor in technical progress. One of the early advocates of this form of enlightened self-interest was Dr. J ames Douglas, whose name is honored and revered in mining circles for his intelligent leadership. In an address on Secrecy in the A7"ts he not only encouraged publication and exchange of ideas, but regarded secrecy as a definite barrier to progress. Pointing to the decline of Swansea, Wales, as a great copper-smelting center, he said: "At Swansea, every gate to the smelting-works is guarded, and as a result it has been as difficult for ignorance to escape out as for suggestions to find their way in." If this book makes it possible "for suggestions to find their way in" wherever cyanidation is practiced, its purpose will have been accomplished and its authors richly rewarded.

H. C. PARMELEE

Preface to the Second Edition The revision of Cyanidation and Concentration of Gold 'and Silver Ores was undertaken with two principal objects in view; the first being to bring . the account of gold and silver metallurgy up to date in accord with the considerable technical advances in recent years; the second being to enlarge the scope of the book to include new material relating particularly to the treatment of refractory ores and the chemistry of cyanide solutions which we believe will prove to be of special reference value to both consultant and operator. oJ It also seemed advisable to rearrange chapter headings and contents in order to give the reader a more systematic approach to the text as a whole. The trend, for instance, during the last ten years or so toward treatment schemes which make a greater use of concentration before cyanidation has necessitated a reexamination of this phase of the subject with the placing of greater emphasis on gravity concentration, flotation, and amalgamation practice. Part I of the book deals with the technical aspects of the subject, including only such descriptive material as is necessary for illustrative purposes; while in Part II there will be found descriptive details of a number of the more important and typical operations throughout the world, which in the case of the treatment of gold ores is handled under four distinct global areas. The fewer, but no less important, instances of silver ore treatment are covered in a separate chapter. In addition to the above changes, all details relating to analytical methods have been gathered together under one title and are now presented as an appendix where it is felt they will be more readily available for reference purposes. We wish again to acknowledge the valued assistance of our associates in the preparation of this revised edition. Particularly are we appreciative of Mr. A. D. Marriott's contribution on South African methods, Mr. C. Blackett's notes on recent advances in Australia, and of the help of the many mining company officials in various parts of the world who cooperated so willingly in supplying information on current practice.

J. V. N. DORR F. L. BosQUI WESTPORT, CONN.

May, 1950

ix

Preface to the First Edition. My purpose in writing this book is to record current cyanide practice throughout the world, giving only enough historical facts to serve as a background for the present development of the process and of the equipment used in its application. The inclusion of general testing procedure and numerous performance data is designed to ensure maximum usefulness to the reader. ., The fundamentals of mill practice and general types of equipment have changed little in the past fifteen years, although individual machines and general technique have been greatly improved. Outstanding developments have been confined principally to crushing, grinding and classification, adoption of flotation for certain types of ore, and precipitation practice. A few of the older and well-designed plants not now in operation have been described because their practice was not far different from the best today, and because quite complete data on their operation were available. The description of plants now operating, both old and new, gives a fairly broad cross section of various practices since the general adoption of continuous methods. I have approached this task with some hesitancy, which I am sure will be readily understood. But my relations with the metallurgical industry have extended over so many years and have received the cordial recognition of so many friends that I feel free to write as if I were still an operator, or consultant only. My first introduction to cyanidation dates back 37 years when I was chemist, and later operator, under lease, at Deadwood, South Dakota, of one of the oldest cyanide mills in America, designed by the Gold and Silver Extraction Company about 1894. The first mill I built and operated, the Lundberg, Dorr and Wilson, at Terry, neal' by, was turned into a profitable undertaking by my invention of the Dorr classifier; and in remodeling another mill in the same district the Dorr thickener was born. The Black Hills district of South Dakota, though small, presented many diverse metallurgical problems and yielded some important developments. Out of it came the first crushing in cyanide solution in America, continuous decantation with mechanical thickeners, mechanical classification, continuous zinc-dust precipitation and the sluicing filter press of Merrill, and the first successful use of Moore's vacuum filter. So many have contributed to the development of the art of cyanidation that it is difficult, if not impossible, to make the record complete. I wish to express my thanks and appreciation to those managers and metallurgists all over the world who have published the results of their work and have xi

xii

PREFACE

so generously responded to requests for technical information; also to those other producers of the tools of the industry who have given their data so freely. I am greatly indebted to my associates for assistance in the preparation of this work, especially to Mr. E. R. Ramsey, "without whose aid it would never have been undertaken, and to Mr. Anthony Anable. Thanks are due also to Messrs. M. VV. von Bernewitz, Burr A. Robinson, and J. C. Williams for their aid in collecting and preparing the material for publication and to Messrs. H. A. Megraw and J. A. Baker for helpful comment and suggestions. JOHN NEW YORK CITY

October, 1936

V. N.

DORR

Contents Foreword , .. " . .. . .. Preface to the Second Edition ' ", . . Preface to the First Edition. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . Abbreviations. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Part I

I II

III IV V VI

VII VIII IX X

XI XII XIII XIV

IX Xl

xv

Technical

Historical ..>....................... Examination and Testing of Ore. . . . . . . . . . . . . . . . . . . . . . . . . . Coarse Crushing. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Sorting and Sampli.ig . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Fine Grinding. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Classification........................................... Sand Treatment. . .. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Slime Treatment..... . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Concentration Roasting............................................... Amalgamation and Bt.llion Recovery. . . . . . . . . . . . . . . . . . . . .. Plant Control. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Cyanicides and Refractory Ores. . . . . . . . . . . . . . . . . . . . . . . . .. Cyanide Regeneration and Miscellaneous Processes. . . . . . . .. Part II

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3 16 46 56 62 74 87 96 128 158 177 209 238 253

Descriptive

XV Treatment of Gold Ores

XVI XVII

. . . . . . . . . . . . .. 279 _. . . . . . . . . . . . . . . . . . . .. 281 Section 1. North America Section 2. Central and South America. . . . . . . . . . . . . . . . . .. 336 Section 3. Africa 354 Section 4. Australia, Pacific Area, and Asia. . . . . . . . . . . . .. 396 Treatment of Silver Ores 428 Costs and Power. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. 446

Appendix A Useful Reference Information " Books on Cyanidation. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Appendix B Analytical 111ethods. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. I rulex . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . ..

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457 484 487 503

Abbreviations In abstracting from the literature a certain style of abbreviation has been maintained, as follows: Bulletin of Canadian Institution of Mining and Metallurgy (Bul. C.I.M. and lYI.) , Montreal, Canada. Bulletin of Institution of Mining and Metallurgy iBul, I.M, and 111,), London. Bulletin, Injormaiioti Circular, Report of Investigation, 01' Technical Paper' of United States Bureau of Mines iBui. I,C" R.I., T.P.U,S,B. of M.). Bulletin, Professional Po/per, of United States Geological Survey iBui., P.P., U.S.G,S,). Canadian J1,lining Journal (C,M,J.), Gardendale, Quebec. Engineering and Mining Journal (E. and M.J,), New York City. J ournol of Chemical, Metallurgical and Mining Society of South Africa (Jour. C.l11. and 1l1.S.S,A.), Johannesburg, Transvaal. Mining and MetalluTgy (M, and M.), New York City. Mining and Scientific Press (M. and S.P.), San Francisco, California. Mining Magazine (M.M,), London, England. Mining Journal (M,J.), London, England. Proceedings, Australasian Institution of Mining and Metallurgy (Proc., A.I.M. and M,), Melbourne, Australia, Proceedinqs, Institution of Mining and Metallurgy (Proc. I,M. and M.), London, England. South African Mining & Engineering Journal (S,A,M. & E.J.). Transactions, American Institute of Mining and Metallurgical Engineers (Trans., A.I,M.E.), New York City.

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Part I

Technical

I

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CHAPTER I

Historical The process that was to revolutionize all concepts of gold milling and increase a gold output that was even then lagging behind the needs of an . expanding world-credit structure was developed without financial backing by three zealous experimenters in a poorly equipped laboratory in Glasgow, Scotland, almost half a century ago. To J. S. Mae.Arthur, a metallurgical chemist, and to R. W. Forrest and W. Forrest, doctors of medicine, the cyanide process for gold extraction owes its origin. At the time of their inadvertent discovery, MacArthm was chief chemist to the Tharsis Sulphur and Copper Co. in Glasgow, where he had experimented extensively on the recovery of small quantities of precious metals from copper leaching solutions and on increasing recovery by the use of chlorine and bromine. As this was outside his company's usual activities, a small syndicate was formed, including Mac.Arthur, the Doctors Forrest, and George Morton, a Glasgow businessman. A room at the Forrests' surgery was fitted up as a laboratory. The results of some of his work were published by MacArthur in the Journal oj The Society oj Chemical I nclustry, and as a result he was asked to investigate the Cassel process for the Cassel Gold Extraction Co., the inventor of the process, H. R. Cassel, having unexpecteclly resigned. While Macarthur reported that the Cassel process for gold extraction, depending on the solvent action of electrolytically generated chlorine, could never become an economical process, he continued experiments in his laboratory on other solvents, with the result that the dissolving action of a dilute solution of cyanide on gold was discovered. Although L. Elsner, aGerman chemist, had published the findings of his experiments, which included the basic idea of cyaniding, in J Ol~r. f. pra7ct. Chem. in 1846, he had failed to recognize its significance and made no practical use of it. It was MacArthm and the Forrests who realized the importance of their discoveries and fathered the idea through the vicissitudes of lengthy litigation, demonstration and finally commercial realization. On Oct. 19, 1887, they registered their first patent-British Patent 14,174. It covered the efficacy of potassium cyanide as a solvent in weak solutions. Their second patent, registered the following year, included the use of alkalis, methods of applying cyanide, and the use of zinc "in a state of fine subdivision" for precipitation. A year later their extraction and precipitation inventions were patented in the United States by U.S. Patent 403,202, 3

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CYANIDATION AND CONCENTRATION OF ORES

granted on May 14, 1889. Practically without exception the patentees were unsuccessful in upholding their claims in the face of determined litigation in the Transvaal, New South Wales, Tasmania and finally in the United States. The original claims of MacArthur and the Forrests are remarkable in that little has been disproved. Their claims with respect to the alkalinity of solutions (weak ones preferred) and to the use of zinc as a precipitant were basic. Nevertheless, the inventors were forced to modify their initial claim regarding the strength of solution and also that in connection with the use of finely divided zinc as a precipitant was circumvented, in the early days, by the use of zinc filament or zinc shavings. Following the growth of the cyanide process in the hands of other men, we trace its progress in the world's major gold fields-Australia, New Zealand, the Transvaal and the United States-for by 1888 in each of these areas British cyanide specialists were attracting attention. The first commercial cyanide plant was erected at the Crown mine, at Karangahake, New Zealand, in 1889, by J. McConnell. Alfred James, representing the MacArthur-Forrest company, introduced cyaniding on the Rand in 1890 at the Robinson mine. The first cyanide plants appeared in the United States in 1891, one at the Consolidated Mercur in Utah, designed by H. W. Brown, and the other in Calumet, Calif., managed by A. B. Paul. Mexico's first cyanide plant was at El Oro mine, of the American Mining Company, in 1894. South African statistics show the rapidity with which the cyanide process was accepted in its infancy, for, from $6,000 in 1890-the first year of cyanidation-the value credited to the process increased to $6,000,000 in 1893. In the early days in New Zealand and Australia the stamp battery was in the period of its greatest use, dry crushing hand-sorted ores. Crushing was followed by roasting and chlorination, since the ores were complex and not amenable to simple amalgamation. With the advent of cyanidation the chlorination process, a costly system, declined and never revived. Later, as the metallurgy developed, weak cyanide solutions were used in the crushing systems instead of water. Sand and slime were separated, first in pointed boxes or cones, the sand going directly to leaching vats equipped with filter bottoms, where leaching baths were applied in succession, as required. The slime was settled in tanks and agitated in cyanide solution. Mechanical agitation was first used alone; then compressed air was introduced through tubes reaching to the bottom of the tanks and later by air lifts in tall cone-bottomed tanks-the Brown or Pachuca tank and the Parral tank, a modification of Brown's air-lift idea-and finally the Dorr agitator, combining air and mechanical agitation. Slime thickening and washing were accomplished in settling

HISTORICAL

5

tanks equipped for the decantation of clear, supernatant solution or in filter presses provided with facilities for washing the cake. Some' ores were roasted before cyaniding, and others were treated successfully by the bromocyanide p r o c e s s . ' . Generally, fine zinc thread, or shaving, was used for precipitating-the precious metal from cyanide solutions, employing upward percolation in the well-known baffle-type zinc box. Occasionally charcoal was used as.a precipitant, and in Africa electrolytic precipitation on lead-foil cathodes. The' application of the cyanide process to ores the value of which is principally or wholly in their silver content lagged behind its use with gold ores. This is largely because silver usually occurred in combined form, sulphides and chlorides. It was believecl at first that these compounds would not yield to cyanide treatment. About 1900, however, Leonard Holms made preliminary tests upon the silver sulphide ores of the Sirena mine at Guanajuato, Mexico, and soon thereafter E. M. Hamilton built and operated a large-scale experimental plant at the same property, definitely proving the usefulness of the process with ores of that character. Previously, Hamilton had operated the plant built by Charles Butters, at Minas Prietas, Sonora, treating by cyanide a large quantity of goldsilver tailings accumulated from the operation of the Grand Central panamalgamation mill. Butters and Mein in 1894 devised a revolving-arm distributor for uniformly charging vats, which made for economy and promoted more uniform leaching. In 1895 H. L. Sulman and F. L. Teed were working toward improved precipitation methods and probably were responsible for the stimulus that led up to the perfection of zinc-dust precipitation methods by C. W. Merrill at Marysville, Mont., and later at the Homestake. Caldecott is credited with having developed agitation in 1898 from mechanical to air methods, starting the sequence that led up to the most modern methods. In April, 1901, Blaisdell patented his excavator for plowing leached sand to a central discharge outlet, whence it was carried by belt conveyor to waste. About 1899 L. H. Diehl introduced the tube mill, to accomplish finer grinding, borrowing the device from another industry. While progress in cyanidation continued in Australia and the Transvaal, it was in the newer fields of Mexico, the United States, and Canada that the greater number of improvements in the art had their genesis. During my work in cyanidation in the western United States between 1899 and 1912, as chemist, as cyanide mill owner and operator, and finally as consulting engineer and plant designer, direct needs and an early recognition of the importance of continuous, foolproof operation, as well !1S the

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fundamental advantages of mechanical control of liquid-solid mixtures, led me to the invention of the DOlT classifier in 1904, the DOlT continuous thickener in 1905, and the DOlT agitator combining air and mechanical agitation in 1907, as well as to development of continuous countercurrent decantation. While each was produced to meet an individual need, all contributed materially toward higher recovery, lower costs, and larger units by making finer grinding and a single all-sliming process more feasible. Many have contributed to improvements on these basic inventions. A. L. Blomfield, a former associate in a consulting capacity, developed a type of tray thickener that has been widely used and later the bowl-type classifier. Both machines were developed at the Golden Cycle mill, of which he was manager, to meet specific demands of mill expansion. About 1903, George Moore, closely followed by Charles Butters, was applying the principle of vacuum-leaf filtration to cyaniding, and their filters came into wide use. The invention of E. L. Oliver, in 1907, of the segmental rotating vacuum filter made continuous filtration and washing possible and tended to supersede filters of the Moore and Butters types. Now the Oliver and other types of continuous filters are used in all parts of the world. . During the same period, continuous precipitation using zinc dust, excluding air from the system, was developed by C. W. Merrill, and this was perfected by T. B. Crowe, who applied vacuum to remove air from the solutions before precipitation. Prior to the war period, 1914-1918, processes of cyanidation had become crystallized, the forerunners of present equipment had appeared in more or less developed forms and the industry was ready for consolidation of the advances. Thereafter, refinement of details received more attention, and the cyanide process entered its present stage. Gold production, which amounted to 21,303,725 oz. in 1914, was increased by discoveries, developments and new practices to a total of 27,474,516 oz. in 1934, the record production up to that time. Multiple-stage crushing became standard practice, finer stages using rolls or cone crushers in closed circuit with screens, generally of the vibrating type. Larger jaw and gyratory crushers were used to hanclle larger pieces of ore, and fine crushers of improved design made possible a product of smaller size, 74: in. or less. The stamp battery, still used in South Africa and a few mills in North America, was operated principally as a finecrushing machine. Multiple-stage fine grinding is now largely preferred with the grinding machines in closed circuit with mechanical classifiers. This latter period has brought emphasis upon the effort to reduce power costs and at the same time to grind finer than had been considered practically possible. The heavy-duty ball mill, short in proportion to its di-

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7

ameter, loaded with steel balls, has become of great importance. The advantage of high circulating loads between grinding mill and classifier led to the introduction of heavy-duty classifiers of both single stage and bowl type, and this trend is increasing. Many mills are producing a final pulp containing only a small percentage of plus 325-mesh solids. In 1925 J. J. Denny, at the McIntyre Porcupine mill in Ontario, iiltroduced differential grinding and selective agitation for the McIntyre ore, in which a large portion of the precious metals was locked in a refractory pyritic envelope. A bowl classifier at the end of the agitator series effected a concentration of the heavy pyrite, in the rake product,which was returned to the head of the mill for regrinding, while the less valuable quartz overflowed to decantation direct. Not only was the pyritic constituent ground finer than the quartz, but also it remained in the agitators much longer. Thus the quartz and pyrite were each ground and agitated to the degree demanded by their relative values and amenability, yet no uneconomical overgrind occurred. Much the same thing was done in South Africa at the Spring mines a little later. In the new McIntyre mill, built in 1931; Denny introduced flotation cells placed between ball mills and the classifier to extract concentrates from the system as early as feasible. Vacuum filters developed rapidly, and their use spread into all camps in the 1920's, especially those of the rotating-drum type of Oliver and, in lesser degree, the American disk-type filter. The first drum filter in Canada was used at the Hollinger mill for dewatering and washing the tailings from its countercurrent decantation plant. At first, vacuum filters were used only for dewatering tailings or, at most, one stage of washing, in which case barren solution or water was sprayed on the cake to displace gold-bearing solution. Later, there was introduced a new system known as double filtration, wherein two filters with an intermediate repulping agitator were arranged in series for washing the pulp by a combination of spray washing and displacement. The question of how many stages of decantation in thickeners and how many stages of vacuum filtration to use is dependent upon economic conditions, and while no definite trend is discernible, it appears that average practice favors three to four stages of decantation, followed by one stage of vacuum filtration. South African conditions are unique because Rand ore filters readily, and the slime can be washed easily by one stage of filtration. Intermittent filters of the Butters type were used to an appreciable extent, but new plants are now installing continuous rotating filters, A study of cyanide methods indicates that no system has become more universally used than the Merrill-Crewe precipitation process. The use of zinc shavings has practically disappeared, except in the older mills. Clarificationof pregnant solution in presses, deoxidation in vacuum towers

8


(Crowe process), addition of zinc dust under conditions precluding the reabsorption of oxygen, and, finally, the collection of the precipitate and excess zinc dust on filter presses or in filter bags have become standard practice. Potassium cyanide was used generally in the early days of the cyanide process but was replaced later by the cheaper sodium cyanide. In 1917, owing to an acute shortage of white cyanide resulting from war conditions, black or "Aero" cyanide, Ca(CNh was introduced. Real del Monte, in Mexico, was the first to use the new cyanide in 1917, and such representative mines as Golden Cycle, North Star, and Tonopah Extension rapidly followed the example of the Mexican mine. Today, black cyanide is extensively used in every mining district in the world. Of great interest in precious-metal metallurgy is the flotation process long practiced in copper, lead, and zinc concentrators. Its use in gold and silver milling really goes back only to about 1931, although countless laboratory and semicommercial scale tests antedated its use on a full-plant scale by more than a decade. Sometimes it is used alone for the recovery of gold and silver, yielding a concentrate which is shipped to a smelter and a tailing which is discarded, but usually it is employed in conjunction with cyaniding. Several options, in this latter regard, are available, such as flotation of the mill heads, followed by cyaniding the concentrates or cyaniding the heads with retreatment of the tails by flotation. Flotation, to be sure, is not a solution for all milling problems, for its use is limited to certain ores, but it has become an increasingly valuable tool for the metallurgist. In the course of its development the cyanide process has made material contributions to many forms of technology. TIllS applies not only to basemetal metallurgy but to nearly all industries where the handling of finely divided solids in liquids is involved. The early recognition by those who developed cyanidation of the importance of mechanical control, the continuous treatment of fine solid-liquid mixtures, and the use of large-scale units has opened the way for advances in sewage treatment, water purification, and many of our wet chemical and industrial processes. WORLD'S GOLD PRODUCTION'

Owing to the gradual depletion of known deposits that were profitable to work at costs existing prior to 1929, the world's gold-mining industry was slowly but steadily shrinking in importance. In 1928 the output was less than 20 million ounces. As John J. Croston pointed out in 19361 (and this is equally true today), the decline in production was based upon the inexorable fact that the deeper 1

"Effect of Revaluation on the Gold Mining Industry," T.P. 709, A.I.M.E.

HISTORICAL

9

you dig and the leaner the ore treated the higher become the production costs. Several countries offered a bonus on gold production as a means of helping the struggling producers, and this, in effect, was the same as revaluing the gold as far as its sale by the mines was concerned. Croston further comments: If the world wants gold in sufficient needs for world commerce, it will have to pay a price commensurate with the cost of producing it. Gold increased almost four times in price during the period 1344-1717, when it was pegged at the equivalent of $20.67 per troy ounce. It would be ridiculous to attempt to prove that this price bore any relation to the cost of producing gold either in 1717, in 1890 or 1929. It was purely an arbitrary figure but with the discovery of rich deposits in California, Australia, and South Africa was sufficient to bring forth a requisite volume of production. Methods of mining and treatment have reached a high degree of efficiency, and few radical reductions in per-ton costs are to be expected. The paucity of new discoveries and the necessity for treating lower grades from greater depths indicated a gradual drying up of production unless some stimulus were given to the industry.' This gradual shifting of the economic phases of world gold production has seemingly escaped the attention of the majority of monetary writers. They have treated gold production as a static affair unaffected by changes in reserves, grade, depth, and other factors in exploiting this wasting natural asset. Certain phases of the economics of gold mining are not so simple as those of hog raising or wheat growing.

REVALUATION OF GOLD

The price of gold in the United States, formerly fixed at $20.67 per ounce, began to rise on Apr. 20, 1933, and continued to rise until it gradually attained $30 per ounce during September of that same year. It was fixed at $35 per ounce by Presidential proclamation on Jan. 31, 1934, a price that was maintained through 1935 and to date, the first part of 1950. The price of gold in other world markets approximated the United States price range. The steady rise in the world's gold production after 1929 continued up to a peak of 42 million ounces in 1940. Thereafter, during and following the Second World War, the production declined as steadily as it had increased in the earlier period. The initial causes of this decline were, of course, the diversion of men and materials to the war effort and such emergency measures as the W.P.B. Order L-208 which suspended most gold'The first fillip to gold mining came with the onset of the depression of 1929-1935. Labor became more abundant and efficient; wage rates dropped, as did the cost of supplies, enabling marginal producers to gain a better foothold. The second boost came when Great Britain went off the gold standard, enabling the producers in certain countries to benefit by a substantial exchange premium. The last and greatest step was taken when The United States government officially revalued gold from $20.67 to $35 per ounce. A careful consideration of the background factors leads to the conclusion that gold will not go back to its old price.

f-'

o

(")

>-1 ~

Z H

tj ~

>-'l H

o Z

~

ztj o

OOOHAWAIl

o

Z

o

i:'I

Z

>-'l

l:O

~

>-'l

H

o

Z

o

'cI

o

l:O

t;=j U2

FIG. 1. Gold-producing centers of the world, according to S. D. Strauss in E. and M. J.

are the producers of most of the silver ores.

Canada, United States, and Mexico

HISTORICAL

11

mining operations in the United States. War damage in such areas as the Philippine Islands and New Guinea was also a contributing factor. But even with the cessation of hostilities production continued at a low level due principally to rising costs; higher taxation, and the fact- that the producers were under these circumstances forced to market a. commodity at a . fixed price, irrespective of the difficulties of acquisition. WORLD'S SILVER PRODUCTION

Silver production has for the most part followed the same course as gold production and to a large extent from the same causes. Reaching an alltime high of 277 million ounces in 1937, the Clecline which started in 1941 had not been arrested by 1945. The price of bar silver for 1939 (London quotation based on current rate of exchange) was $0.408 per ounce. The United States government price for newly mined domestic silver, which had been $0.646 up to July of that year, was increased on July 6 to $0.711 per ounce. At the beginning of 1945 domestic consumers obtained their silver from sources completely under government control. After July 1, 1946, the price for domestic metal was, however, again raised to $0.905 per ounce, where it remained for the rest of the year. Table 1 showing the world production of gold and silver in 1946 was prepared largely from the latest available data, which, however, were stated to be in many cases incomplete. It does show, nevertheless, the approximate production rating of the various countries and draws attention to the interesting contrast between the magnitude of the gold and silver output for individual countries. ECONOMIC ASPECTS OF GOLD PRODUCTION

Because of the disturbed economic conditions following the Second World War and the depreciated value of most currencies in terms of buying power throughout the world, a number of articles and reviews on the subject of the monetary aspects of gold have appeared in recent technical publications. Writing in E. and M.J., 1. B. Joralemon" reminds us that ... gold has been the symbol of wealth in all the civilizations that have arisen in the past 10,000 years. The Assyrians; the Egyptian dynasties; the ancient empires of India and China, Greece and Rome; the Incas and Aztecs; and all the other peoples that have made a brave show on the pages of history have worshipped at the feet of the Golden Calf. In modern times the hopeless depreciation of paper currencies has turned men back to gold for security wherever governmental action has not made the holding of gold a crime. Throughout history men have toiled and fought for this beautiful and enduring metal. 3

"Who Has the Gold," E. and M.J., Vol. 149, No.7, p. 76.

12

CYANIDNrION AND CONCENTRATION OF ORES

In spite of the tremendous influence of gold on mankind, for good and for bad, all the gold that men have won from the earth in 10,000 years could be put in a room 48 by 40 by 20 ft. Less than a seventh of this gold has been destroyed or lost.

Estimating the total world production during this period as 1,616 million ounces, Mr. Joralemon deduces that the nations and Central banks now TABLE

1.

WORLD GOLD AND SILVER PRODUCTION,'"

Gold Country

1946

Silver

Troy oz.

11,917,914 6,500,000t 2,807,643 1,462,354 824,480 587,000 544,596 437,176 420,500 331,394 230,521 200,000t 181,615 158,378 131,680 119,271 77,931 73,000 65,100 706,447t 27,777,000

Country

Troy oz.

43,263,132 21,103,269 12,676,928 12,334,761 1O,000,000t 9,073,481 6,106,165 5,047,666 2,682,910 1,203,978 1,107,827 1,000,000t 698,636 313,180 223,308 192,200 152,651 127,222 95,168 1,597,518t 129,000,000

'" Compiled from data given in Minerals Year BODle, U.S. Department of the Interior, 1946, and other sources. t Estimate.

hold 61 per cent of the total and individuals and corporations 25 per cent, leaving only 14 per cent unaccounted for since man first started mining. No other human possession in all time has been so zealously and effectively guarded. Who has the gold? We in America have most of it, through our share in the great government reserves or in gold ornaments. Outside the United States those who can afford to own gold hold it as a favorite method of insurance against future want. Throughout the ages, man's faith in gold has never wavered, and gold has never failed him. Men cannot eat gold, but those who have had gold have never gone hungry.

HISTORICAL

13

Disturbed by the haphazard functioning of currency relationships, a group of members of the A.I.M.E. San Francisco Local Section and of the Mining and Metallurgical Society of America have undertaken a scientific study of the currency problems.' It is their contention that difficulty of acquisition is the -real measure of value of both gold and currency and that it is possible to express the r'elationship between them in mathematical form. Reasoning along these lines, they reach the interesting conclusion that the true currency price of gold was, for instance, $65.73 per ounce in 1945. The question as to whether the resumption of a free market in gold would be compatible with public interest is discussedby the Empire Trust Company of New York in their Letter 44 of Mar. 4,1949. Quoting Professor Kemmerer" as "probably the foremost authority on the gold standard," J. S. Lawrence, vice-president of the company, says: Although there are many types of the gold standard, the gold standard may be said to exist in any country in which prices of goods and the obligations of debtors are usually expressed in terms of the value of a monetary unit consisting of a fixed quantity of gold in a free market, The gold standard exists whenever the value of gold in a free market is the actual standard, regardless of the machinery by which the standard is maintained and regardless of whether this machinery operates automatically or is managed.

Again we are reminded that a free market for gold is a form of free speech. It gives the voter, the trader, the speculator a chance to pass judgment on the fiscal conduct of his government. In view of all that has happened since 1934, a free gold market is the necessary prelude to an honest gold standard. It is the only practical way to restore such a standard. It is proposed that, since no one can know what a proper basis for the return to the gold standard should be, the practical way to find out is to repeal the penalties on the possession of gold, permit all gold currently produced or imported to be traded in an open market, and so allow it to find its natural price level. It would then be safe for the government to open the doors of its mints and redeem at the statutory rate wherever paper currency is offered. In a paper presented at the annual meeting of the A.I.M.E., New York, February, 1948, D. H. McLaughlin 6 says: Prices paid for goods and services in paper currencies are undoubtedly determined by many interrelated factors, but among them none is more specific in pushing ,j "What Price Gold? A Scientific Approach to Finance," M. and M., January, 1948. 5 E. W. Kemmerer, "Currency Stabilization in Latin America," Fourth Pan American Commercial Conference, p. 2, Oct. 6, 1931. 6 "Gold versus Inflation," M. and M., April, 1948.

14

CYANIDN.rION AND CONCENTRA'l'ION OF ORES

prices toward higher and higher levels than the vast increase in debts and monetary obligations of our own and prncticnlly all major governments. There is nothing unorthodox in the situation in which the gold miners find themselves today. With wages and the prices of 1111 materials required in the production of gold I1t the high levels ohnrnoteriatio of a period of prosperity, phony or otherwise, and with shortages of both men and machinery to contend with as well, the spread between the returns in paper dollars that must be accepted for gold in the United Stutes and the cost of producing it is ruther painfully small for most producers. Technical improvements, stimulated by the urgency of declining profits, have to some extent made the decline in profits less drastic in 11 few cases, but the relief to be gained in this WI1Y in an industry already smoothly and efficiently operat.ing is I1t best rather limited. Consequently we find the gold mines the country over still far below the level of output and profits that formerly prevailed arid still struggling to overcome the special financial and physical hardships that were so nrbitrarily imposed on them during the war. On the other hand, faith in the inherent value of gold seems stronger than ever, even though the returns from mining are temporarily below the rate that makes the industry I1S 11whole prosperous. This faith is revealed by the persistence of prospecting and exploration in 1111 the gold-bearing regions without apparent decline in enthusiasm and hope, and it is most clearly shown by the bold and immense expenditures on deep shafts and large plants that are being made with confidence in new enterprises on the extensions of the gold-bearing reefs of the Rand and on the newly discovered deposits in the Orange Free State. It would be most interesting to know at what price per ounce gold would circulate freely with today's paper dollar. If such a price could be established by some practical and empirical means, stabilizution of the dollar I1tthat level with complete interchangeability of gold and pl1per would undoubtedly be a most beneficial step. With this accomplished, adjustment of the various currencies of the world to gold and the dollar would be 11 relatively easy way of correcting the fractions that are solemnly agreed upon in today's official exchange rates. A higher price of gold would no more influence domestic prices in dollars than did the increase from $20.67 to $35 per ounce in 1934. But it would be an admission of the weakness of the dollar and of the present degree of inflation,

In a paper entitled "The Future of Gold," Arthur Notman? remarks that apparently, the world as a whole still regards it (gold) as preferable to the paper promises of governments, however sound, relatively, they may appear. Actually, we have no way of measuring the true strength of this feeling, because the principal governments of the world have made it legally impossible for their citizens to own gold except in the form of ornaments. If the volume of gold in the world fails to grow in step with the volume of the world's business, because it has become more difficult to find and produce I1t I1n arbitrarily set price in terms of other goods or because of an increasing demand for the metal as a means of conserving values (through hoarding), or because of the growing volume of paper promises, thereby reducing the availahility of the metal for free conversion, there remains but one way in which to restore that confidence 7

Trans. 50, eLM. and M., 127-130, 1947.

HISTORICAL

15

necessary to trude, namely, by facing reality in recognizing the decline in the purchasing power of paper currencies through rnising the value of gold in terms of those depreciated tokens and encouraging the search for and production of gold and its release from hoarding. If we are not to return to the "dark ages," which were "dark" primarily because the world's confidence in the integrity of its neighbours had disappeared and with it the free exchange of goods, services, and ideas had come to a halt, we must brea"k this vicious circle. It has been suggested in certain well-informed quarters that the first step in this direction should be the restoration of free coinage of gold by the United States, as the only nation in a position to do so. The United States dollar would then cease to be merely another paper token, and other currencies could be accurately weighed against the dollar. Many of the uncertainties in the proposed operations of the International Bank and Fund wouldfbe solved by this procedure, in their opinion. With little or no claim to wisdom in the field of monetary policy, I very humbly record my endorsement of this procedure.

Addressing the 75th Anniversary Jubilee of the A.I.M.E. on "The Future of Gold in the World Economy," P. M. Anderson," managing director of Union Corporation, Johannesburg, saw no basis for fear over gold's future status. He viewed gold as having a function in monetary systems similar to that of stand-by capacity in a power plant and predicted that democratic nations will return to gold as rapidly as they can solve the emergency problems which have grown out of the war. Presuming that the Far East will improve its economic status, he saw it as a market for nearly half the gold production of the future. He predicted that the financially stronger democratic countries would soon make gold available to citizens and that the financially weaker ones would find gold a necessary medium of exchange because of lack of confidence in their paper. Mr. Anderson concludes by remarking: At the present price of the metal, gold will be produced in quantities sufficient for a steady expansion, but not in any great abundance. Gold will presumably be added to the world's stock at a rate closer to 2 per cent than to 3 per cent per annum. 8

E. and M.J., Vol. 148, No.4, April, 1947.

CHAPTER II

Examination and Testing

0/ Ore

Repuioble firms and experienced metallurgists who specialize in the testing of ores are to be [tnuul in all of the principal mining countries. Laboratories operated by governmental agencies conduct preliminary tests on ores for prospector 01' mineowner in the United States, Australia, and Canada. Leading mamufaciure» of machinery and chemicals used in cyanide and flotation plants maintain testing plants where the essential factors governing the proper llse of their products are tleiermined by experienced experimenters. For all those properly qualified, by education 01' experience who desire to carry out their own tests, a description of the general methods of conducting them and of the a/pptmuus required. is O1ltlined in this chapter. SAMPLING

The value of any series of experiments depends entirely upon the kind of sample of ore provided for the test. It must represent, wholly and completely, the character of the material that will be treated in the plant to be designed and constructed. If it does not, the tests will be useless or, worse, misleading. In some cases, owing to conditions under which mining is carried on, it is considered advisable to remove a portion of valueless material by sorting before the ore is delivered to the metallurgical plant. No intelligent appraisal of the need for sorting or its extent can be determined from a laboratory sample, but it is essential that the test sample should represent accurately the character of the proposed mill feed, as to both precious-metal content and the proportions of barren material, wall rock, gangue, etc. All these factors are important, since they determine the sizes of equipment for settling, filtering, clarifying, and classifying as well as grinding equipment. These details are crucial and are essential parts of a complete investigation of possible metallurgical processes. Therefore, if sorting is to be done, the test sample should represent the ore after sorting. The quantity of the sample is important in its bearing on' the representative nature of the material. It should be sufficient to represent truly the character of the ore. Canadian custom at present is to ship from 500 lb. to 1 ton for Ottawa tests. Where the material is unusually uniform, smaller amounts will serve as well. In cutting down the sample and making separate portions for different methods of treatment, all possible care should be taken to ensure similarity 16

17

EXAMINATION AND TESTING OF ORE

of each portion. No care is too great to make certain of the representative character of the sample for testing. Consideration should be given also the abnormalities of ore feed that may be expected, owing to variations of the ore bodies being mined, to alter the average of usual character of the mill fued. . Various theoretical studies have been made to determine the minimum size of a sample cut for any specified degree of allowable error. Reduced to its simplest terms, one accepted formula- is

75D3 W - -----------..,. - (01 )2 X (m. !lJom) 10 error l Gm where W = weight of sample required in grams D = diameter in millimeters at which sample is 1.5 per cent plus % error = error allowable (relative) wt. %m = weight per cent of mineral m. dm = density of the mineral ni It is on gold ores carrying relatively coarse gold that the real problem arises, because the wt. %m is low and the dm. is high. In ore assaying 0.1 oz. per ton, the wt. %m = 4.2 X 10-4% dm. = 10.3. Thus if such an ore is to be sampled to 0.005 oz. (= 5 per cent error) at 20 mesh, the sample to be taken is

W=

75 X (.833)3

80,000 grams

(5i X (4.2 X 10 4) 19.3

=

176 lb.

TIlls assumes the gold to be free and occurring in grains up to 20 mesh in size. For a 500-gram sample of this ore to be representative to a 5 per cent error, the ore would have to be crushed to 0.833 V160

=

0.156 mm

=

100 mesh approx,

MICROSCOPY AS AN AID IN CYANIDATION PROBLEMS

The assistance that can be obtained from the microscope in solving oredressing problems has been increasingly appreciated in the recent years, as attested by the frequency of papers on the subject. Observations by means of binocular microscopes with magnifications up to 100 diameters and corresponding resolutions have long been common 1 For the development of the above formula see Gaudin, Principles of Minerol Dreseinq, McGraw-Hill, p. 515, 1939.

18


practice, and such observations are a great help in working out methods for the satisfactory treatment of an ore. Further progress resulted in the use of metallographic equipment and magnifications up to 500 diameters. However, it is only recently that full use has been made of metallographic microscopes with useful magnifications up to 1000 diameters wherein particles 0.5 micron in diameter or less are clearly resolved. Such work requires skillful preparation of the specimen before the full capacity of the microscope can be made use of.

(a)

(b)

FIG. 2. Photomicrograph of a polished surface of a cyanidation residue reveals (a) an inclusion of gold (G) measuring approximately 2 by 3 microns, locked within a grain of arsenopyrite (zlr ): (b) two minute gold particles (G) enclosed in a grain of pyrite (Py), the larger of the two inclusions being only about 5 microns in width. Magnification in each case = lOOOX, with oil immersion objective.' (American Cyanamid Co.)

The latter technique is of especial value in connection with cyanidation of gold and silver ores, because in certain ores some of the gold occurs as minute blebs or stringers in pyrite, the inclusions often being as fine as 1 micron in diameter. Two excellent examples of such occurrences published some years ago are (1) "Increasing Gold Recovery from N oranda's Milling Ore," by C. G. McLachlan, Trom« 112, A.I.M.E., 570, 1935, and (2) The Role of the Microscope in Ore Dressing, American Cyanamid Company, pp. 12-14, 1935. By permission of American Cyanamid Company the following illustrative example is quoted from the paper referred to above: Typical Problems-Example I. A sample of auriferous-pyrite concentrate assaying 4.75 oz. per ton was submitted to our ore-dressing laboratory for the purpose of determining whether the gold could be extracted by means of straight cyanidation.


19

Preliminary tests on this sample indicated t.hat this mater'inl W[1S very refrnct.ory. Subsequent tests, in which the raw concentrate W[1S reground to -325 mesh and cy[1nided for a long period of time with a strong solution, failed to improve the extraction mutei-ially. The lowest eyunide residue contained 0.33 011. per ton. At this stage in the invest.igat.ion it W[1S decided to submit a sample of the above residue to the mioroseopical Iaborutory for the purpose of determining the' form and manner of associ[1tion of the gold. Accordingly, samples of the +325-mesh and -325-mesh residues were briquetted, polished, and examined with the metallograph [1t high m[1gnific[1tion .... It W[1S noted that the gold occurred [1S metallic gold completely encased in pyrite and that the size of the gold particles W[1S about 1 to 3 microns. Inasmuch [1S minus 325 mesh is about the present economic limit for grinding, it W[1S useless to proceed with further tests along the lines of finer grinding and J[1rly eyanidution of the r[1W concentrate. Thus at an e stage in the investigation, the intelligent use of the microscope saved much useless work by narrowingdown the lines of uttack and pointing to [1 practicnl solution of the problem. In this case either rousting prior to eyanidatiou or direct smelting of the concentrate W[1S definitely indicated.

For further information on the subject, the reader is referred to the latter paper, which gives an excellent description of a modern microscopic laboratory for such ore-dressing investigations, with practical examples, methods used, and bibliography. An interesting paper by R. E. Head? presented at the February, 1936, meeting of the A.I.M. and M.E., subject, "Physical Characteristics of Gold Lost in Tailings," indicates the possibilities of the microscope in the study of tailings losses. The following is an extract: SURFACE CONTAMINATION

The experience gained by microscopic study of numerous gold ores and tailings has shown that there are pronounced differences in the physical characteristics of the minerals composing them, more especially of the gold. Repeated studies of tailings from flotation and cyanidation of gold ores has established the fact that surface contaminations on gold particles are often directly responsible for high gold losses; comparison of gold particles isolated from flotation concentmtes and the resulting tailings have shown that the clean gold has been recovered and the tarnished or contaminated gold invariably lost in the tailing. Obviously, it is not possible to make such a comparison of gold ores treated by cyanidation, as the clean gold has been taken into solution; but when the gold found in cyanide tailing has a tarnished or coated surface, one may infer, with a reasonable degree of certainty, that the clean gold has been extracted. This premise is supported by experimental evidence obtained by isolating particles of tarnished gold and exposing them to cyanide solution in small parting cups. In one such experiment, tarnished gold particles picked from a cyanide tailing showed but slight evidence of dissolving at the end of 27 days. In this test, a cyanide solution of 1.6 lb. per ton was used, and the leuch solution was decanted and renewed every 24 hr. The proof of cyanide attack was manifested by a noticeable thinning of the gold particles at the edges. A rim of a substance that appeared to be gelatinous W[1S visible at the edges of the 2 Microscopist, Intermountain Experiment Station, U.S. Bureau of Mines, Salt Lake City, Utah.

20


gold particles, and there is reason to believe that this material encased the entire surface of the gold particles. It is not known whether the gelatinous film is of secondary origin, resulting from a reaction between the cyanide and some substance or substances adhering to the gold surfaces in the form of a coating, or is an original constituent of the surface contamination which has been made visible through the dissolving of a small amount of gold at the margins of the particles.

TESTING FOR ACIDITY OR ALKALINITY

Acidity or alkalinity determinations of water for and solutions used in plants treating gold and silver ores are of importance. Methods range from simple tests with litmus papers or phenolphthalein solutions to pH determinations. pH Determination. The pH value indicates with a high degree of accuracy the amount of active acidity or alkalinity in contrast to the total acidity or alkalinity as determined by ordinary titration methods. The symbol stands for the logarithm of the reciprocal of the hydrogenion concentration H'", and on this account the higher the numerical value on the pH scale the smaller the number of free hydrogen ions in solution, and vice versa. Furthermore, it will be noted that, since the relationship is logarithmic, each successive pH number represents a tenfold change in the hydrogen-ion concentration. Pure de-aerated distilled water is neutral in its chemical reaction, which means that it is neither acid nor alkali and that its number of H ions is equal to its content of OH ions. It has also been determined by analysis that the degree of ionic dissociation of pure water, or, in other words, its hydrogen-ion concentration, is 0.0000001 (one ten-millionth) gram per liter (1000 cc). Therefore by the definition given in the previous paragraph, its pH value would be determined as follows: pH value

=

1+ log (H )

=

log of

0.00~0001

=

7.0

This pH 7.0 value is therefore the neutral point of the pH scale at 22°C. This neutral point rises and falls inversely with the temperature, and at 18°C. it is 7.1. A 0.1 normal solution of hydrochloric acid contains by definition 0.1 gram ionizable hydrogen per liter, so that if completely ionized the solution would contain 0.1 gram hydrogen per liter. From electrical measurements, however, it is known that at '18°C. only 91.4 per cent hydrochloric acid is dissociated into ions. The balance remains in the solution as HCI molecules. Since only 91.4 per cent is ionized, it contains 0.1 X

i~~

=

0.0914 gram H+ per liter


21

Therefore the pH value of O.lN HCI is 1 I og. 0 f 0.0914

or

log 10.94

=

pH 1.04

By a similar calculation it can be shown that a O.lN solution of acetic acid, whose 0.1 gram per liter ionizable hydrogen is dissociated only to' the extent of 1.36 per cent, has a pH of 2.86. In other words, O.lN HCI contains almost 70 times the number of active hydrogen ions as O.lN acetic acid, which by ordinary chemical titration methods is of the same strength. Similarly, 0.3N solutions of sodium bicarbonate (NaHC0 3) , sodium carbonate (Na 2C0 3) and sodium hydroxide ('NaOH) , which all have the same alkalinity when measured by ordinary chemical titration methods, have pH values of the order of 8.40, 11.60, and 13.00, respectively. Because of this difference in ionization, different acids and alkalis are designated as weak or strong. As applied to corrosion, pH values measure the intensity of the corrosive action, while total acidity by titration measures the amount of corrosion which will occur before the acid is exhausted. The method of pH determination is particularly useful where the acidity or alkalinity is so slight as to be below convenient titration range. One method of measurement is to add specific "indicators" which show characteristic color changes at certain pH values and to compare these colors with standards in the form of solution or color charts. One instrument largely used for this purpose is the LaMotte roulette hydrogen-ion comparator, where a large number of standard solutions may be viewed alongside the sample against a fixed light source. Another device makes use of specially treated paper enclosed in plastic containers. Pieces are torn off, immersed in the solution, and the color change compared with charts on the containers (Beckman type). The most accurate device, however, is the pH meter, an electrometric instrument which measures the pH directly against a standard calomel electrode. These instruments, which are made by the Leeds and Northrup Company of Philadelphia, Pa., the National Scientific Laboratories of Pasadena, Calif., and others, are carried by all leading supply houses. SIZING

Sieve Analysis. A sieve analysis of the ore will often supply information of considerable value. This may start at about 20 mesh and include the range of sizes to and including 325 mesh. A portion of each sieve size should be examined under the microscope and assayed for gold and silver. A sample also should be separated into a sand portion and a slime portion. A series of such separations may be made, each starting with the ore crushed

22


to a different degree of fineness. For example, a sample of ore might be crushed so that it all passes a 20-mesh sieve. The material is then separated by panning or by decanting under definite and controlled conditions into sand and slime fractions. Each fraction is weighed and assayed. Another sample of the ore might be ground to pass 65 mesh, and a similar separation made of sand and slime. Yet another separation might be made on the ore ground to pass 150 mesh. In this way the distribution of the precious metals can be determined under different degrees of grinding, and the possibilities for separate treatment of sand and slime thus indicated. As a rule, preliminary cyanide tests are best made by agitation methods alone, and later the possibility of leaching a portion of the ore can be investigated. Elutriation Tests. Because of the mechanical difficulties in making grading tests on sieves finer than 325 mesh, some precise method is desirablewhere finer grinding is being done, as at Kirkland Lake and Beattie (gold) and at N oranda (copper, gold, silver). Trials made with the Nobel elutriator at the last place by C. G. McLachlan are described by him in Trans. 112, A.I.l11.E., 1934. The work was simplified by reason of sizing practically only one mineral-pyrite. At N oranda, they were able to establish five size zones of constant range below 325 mesh with this elutriator. Other types of elutriators are the Schultz and Schoene, which have cylindrical portions at the center, and the Andrews kinetic elutriator, in which a specially designed settling chamber is superimposed on a cylindrical settling column and from which settled sand can be withdrawn into a measuring vessel. A more recent device is the U.S. Bureau of Mines elutriator described in R.I. 2951 and R.I. 3333, U.S.B. oj M. For a more detailed discussion of this subject, the reader is referred to Taggart, Handbook oj Mineral Dressing, Wiley, Sees. 19-99, "Size Testing," 1945. Beaker Decantation. TIns method, which has proved to be reliable and convenient, is in general use at the Westport Laboratory of The Dorr Company. E. J. Roberts describes this procedure as follows: An amount of sample containing about 100 grams minus 325 mesh material is screened on a 200-mesh screen keeping the volume as small as possible. The minus 200-mesh pulp is then washed onto the 325 screen and the material on the screen rinsed well (200 cc water or so) to remove fine slimes. If the material remaining on the screen is over 15 grams, it is divided into 15-gram portions and each portion treated as follows: Into a shallow pan, such as a gold pan, run enough water to submerge the bottom of the screen. Lower the screen into this and jiggle so that the water wells up through the screen. Raise and tap the screen. Continue to do this for 5 min. and at the same time keep swishing the sand around on the screen so that it covers the screen evenly. Dry and weigh the plus 200 and the plus 325. This gives a 44-micron separation.

23


After 1- or 2-min. settling, the water can be decanted from the gold pan and the residue washed into the main body of minus 325-mesh material. This is then allowed to settle if over 2 liters or if much lime is present. The clear supernatant is decanted, and the thick pulp washed into a 2-liter beaker or battery jar. If the sample hus ever been in a lime solution, the safest plan is to dilute up to 2 liters and settle once more to wash out most of the lime. . TABLE 2. Mesh

--Dia., microns ... .

---

12,000 ~ -

1.25 - ~

GOOD

SETTLING RATES OF QUARTZ SPHERES IN WATER* Cm. per min. 3000

1500

-- -- -2.5

5

-- --

325 150 100 750 200 -- -- -- -- --

10

20

--

--

43

74

.1

104

G5 208

147

-- -- -- --

48

35

28

20

-- -- -- -295

-- --

417

595

--

--

833

Temp. "C.

0 5 10

0.281t 0.0188 0.075 0.300 1.200 5.55 16.43 31.8 0.333t 0.0222 0.089 0.355 1.420 6.56 19.45 37.1 0.366t 0.0257 0.103 0.412 1.648 7.62 22.6 42.4

60 107 179 277 422 61o 69 120 196 300 452 652 77 132 212 320 480 69o

0.442t 0.0294 0.118 0.471 1.884 8.71 8.93 0.465t 0.0310 0.124 0.496 1.984 9.16 0.476t 0.0317 0.127 0.508 2.032 9.39 0.488t 0.0326 0.130 0.521 2.084 9.62

229 232 236 239 242

25.8 26.5 27.2 27.8 28.5

49.0 50.1 51.2 52.3 53.4

86 88 89 91 93

20

0.502t 0.0334 0.134 0.535 2.140 9.89 29.4

54.5

94 156 245 365 540 76o

21 23 24 25

0.514t 0.0342 0.137 0.548 2.192 10.13 30.0 0.526t 0.0351 0.140 0.561 2.244 10.37 30.7 0.539t 0.0359 0.144 0.575 2.300 10.62 31.3 0.552t 0.0368 0.147 0.589 2.356 10.88 31. 9 0.564t 0.0376 0.151 0.602 2.408 11.13 32.5

55.5 96 158 56.5 97 160 57.5 99 162 58.5 100 164 59.5 102 166

30

O.632t10. 04210.169

15 16 17 18 19

22

o.453t 0.0302 0.121 0.486 1.932

10.674

2.696 12.46 135.5

64

144 146 148 150 153

248 251 253 256 258

345 349 353 357 361

368 371 374 377 380

510 516 522 528 534

544 548 552 556 560

726 733 74o 747 754

767 773

78o 786 792

109 174 270 395 575 82o 1

* The WESTPORT MILL, DORR COMPANY, INC. t Cm. per hr. For materials other than quartz 325 mesh and finer, multiply above figures by (ds - 1 )/1. 65, where d.s = density of material. NOTE: Multiply em. per min. by 1.97 to get ft. per hr. Divide em. per min. by 30.5 to get ft. per min. The pulp is then diluted back to about 2 liters, 0.1 gram Na 2COa added, followed by 1 to 5 cc sirupy Na2SiOa and the pulp stirred for 5 min. with a high-speed stirrer. Remove stirrer, and if the suspension appears to be dispersed, proceed. Take the temperature, then stir with a rubber-stopper-tipped rod to get as much vertical motion as possible for 10 sec., reverse motion just enough to quench all rotary currents, and note the time. Measure the depth of the suspension from the meniscus to the sand line after tapping. Calculate the time required for a 20-micron particle to settle this distance at the temperature measured by referring to Table 2. Let this = A min. After

24


A min. tap the beaker again, lift carefully, and pour off the supernatant suspension into the other 2-liter beaker. Pour off all the liquid. The +20-micron particles should remain on the bottom of the beaker in a firm cake or at the most only slide a little if the pulp has been properly dispersed. Time can be started on the lO-micron cut as the suspension is poured above. No stirring is ordinarily necessary. The +20-micron material is cleaned while the 'l.O-micron cut is standing. A liter of water is added to the residue in the beaker, 1 cc N a 2Si0 3 added, and the solids thoroughly stirred. Temperature and distance are measured, and the time for a 20-micron particle to settle calculated = A' min. At the end of A I min. the suspension is poured off into a I-liter beaker. This is repeated twice more using the second and third l-Iiter beaker to receive the supernatant. These l-Iiter beakers are placed in line behind beaker 2 which is full of -20-micron pulp. The residual +20-micron sludge is washed into a dish, decanted, dried, and weighed. The lO- and 5-micron cuts are made in exactly the same manner as the 20, the supernatant being poured off at the calculated times for lO and 5 microns. The suspension in the first I-liter beaker is the first wash solution for the lO-micron cut, and it is poured back into the same beaker after the wash, beaker 2 poured on, etc. The wash solutions are suspended thoroughly before pouring onto the sludge so that the beaker drains clean. TABLE

Size of Separation 0.9to1.0 0.8 to 0.9 0.7 to 0.8 0.6 to 0.7 0.5 to 0.6

3.

SIZE OF SEPARATION

Per Cent Settling with Material Coarser than Diameter of Separaiion. 66.4% of the 0.9 to 1. a x material 27.4% of the 0.8 to 0.9 X material 10.0% of the 0.7 to 0.8 X material 5.2% of the 0.6 to 0.7 X material 0.8% of the 0.5 to 0.6 X material

The separations may be carried down to 2.5 or even 1.25 microns if desired, although the times become excessive. Or the butting may be stopped at 10 microns. The overflow from the finest cut is flocculated with A12(S04)3, filtered, dried, and weighed. The percentage of each size is figured on the basis of the total of the weights of the finished samples. . If a 30-, 40-, or 44-micron separation is desired, it is most conveniently made on the finished +20-micron partial, using four I-liter portions of water with no dispersant. The supernatants are all poured into a common receptacle, settled a half hour; the clear water decanted; and the residue dried and weighed as the +20micron fraction. The weight percentages of the micron cuts are reported and plotted at 90 per cent of their nominal diameter; e.q., the per cent in the 20-micron cut is reported as per cent -44 microns +18 microns. The reason for this is that the cuts, even with three washes, are not completely free from finer material but contain enough of it to be just about equivalent to 90 per cent of the diameter. The proportions of the finer sizes contained after four decantations are shown in Table 3.

The rates given in Table 2 are for quartz spheres, density 2.65. As indicated on the table they may be applied to other materials in the fine sizes merely by multiplying by (dB - 1) /1.65.

25


417

_l-

295 208

/

147

1/

104

1/

~ 74

l::?

~

V

/"

52 43

I

?

1/ 18

10 I

2

5

10

II

20 c 40506070 80

-

90 95 9899.5 .8.99

Percent cumulative plus (c)

~

100

E: 80

~

~ 70

L........

60 .~ 50 E: 40 !J>

.~

:§ ~

a

30 ~ 20

I---

.>:

.- V· .i>

10

o

~

0.1

t?

---

I,,~

l--

.02 .03 .04

1.-

.060.8.10

0.2

0.3 0.4

0.6 0.81.0

Opening in millimeters (b)

FIG. 3. Screen analysis of a typical classifier overflow plotted (u) on log probability paper; (b) on log log paper.

Still other methods of subsieve sizing include use of the sedimentation balance" and the Bouyucous hydrometer. These and other related methods are discussed in Mscromeritice by J. M. Dalla.valle, Pitman, 1943. Figures 3a and 3b show two methods of plotting the results of a sizing a Fred C. Bond, "The Sedimentation Balance for Measurement of Size Distribution of Fine Materials," T.P. 1129, A.I.M.E., 1939.

26

CYANIDAUON AND CONCENTRA'l'ION OF ORES

analysis. The first (Fig. 3a) is a plot on log probability paper (No. 3128, Codex Book Co., Inc., Norwood, Mass.) which has the advantage of expanding the readings at the coarse end of the distribution line and makes it possible to compute intermediate values throughout the size range quite accurately. In Fig. 3b4 the same analysis is shown on log log paper with per cent cumulative minus the mesh weight values plotted against millimeter openings. Here the size distribution tends to approach a straight line, and extremely low subsieve readings can be obtained by extrapolation as shown. Air Sizing. Air sizing down to 10 microns, a technique developed by Prof. H. E. T. Haultain of Toronto University, has been established as regular procedure in the testing laboratories of The Lake Shore Gold Mines, Ltd., at Kirkland Lake, Ontario, with most satisfactory results. For details one is referred to the publication of C.I.M. and M., Milling Investigations at Lake Shore, 1936. A brief description of the device is included in Chap. XII for the reader's information. TESTING PROCEDURE

Examination of Ore. Before an ore is tested for its amenability to the cyanide process, it should be subjected to a preliminary examination in order that the experimenter may become familiar with its general physical and chemical characteristics, Knowledge so obtained will be of value in laying out the most effective testing program and may point to special methods of attack should the ore not respondreadily to simple methods of cyaniding. A representative sample is taken from the ore to be tested. The size of this sample will depend somewhat upon the amount of ore available, but it need not exceed ~ to I lb. If the ore is in lumps coarser than 1 in., examination may reveal the nature of the are and gangue minerals, the degree of crystallization, the extent to which crushing may be required to liberate a portion of the ore minerals, and other pertinent information. The sample is then ground to about 20 mesh, and a weighed amount panned. The concentrate from panning is roughly weighed and then examined under a low-power microscope. The amount and size of any elemental gold and silver are noted, and the nature and amount of any sulphides or heavy minerals are determined. Copper minerals and tellurides especially should be sought. A portion of the concentrate (or some of the crude ore) is given a rough qualitative chemical examination. From such a procedure, information as to the desirability or necessity for any treatment such as amalgamation or concentration (gravity or flotation) 4 R. Schuhmann, Jr., "Principles of Comminution. 1. Size Distribution and Surface Calculations," T.P. 1189, A.I.M.E., February, 1940.

EX.AMINATION AND TESTING OF ORE

27

preceding cyaniding may be obtained. Also, there will be some indication as to the degree of grinding necessary. After being ground to a suitable size, the ore is now sampled for assay. This size will depend upon several factors, such as the nature and amount of material available. It is essential that the assay sample be thoroughly representative, and the exact procedure must be adapted to the conditions prevailing. The amount of analytical work done will depend upon the extent of the testing work to be undertaken and will range from determination of gold and silver only to a practically complete analysis. In connection with the analytical work it)s generally advisable to make a qualitative test for water-soluble salts and if any be present to determine their nature. Where the ore is known to contain coarse gold or "metallics," it is essential that these be removed first from the head sample and separately weighed or run down in an assay crucible and collected in lead by standard assay procedure. A known weight of, say, 500 to 1,000 grams is crushed to about 100 mesh and then either panned or put over a laboratory jig for removal of metallic gold. The tailings can then be dried and assayed in the usual way, while the metallic gold recovered is calculated back on the basis of the weight of total sample taken. Preparatory Processes Followed by Cyanidation. The most successful treatment process is the one that yields the greatest net profits. As the complexity of a treatment increases, so also does the number of factors to be considered before arriving at a decision as to the methods best adapted to the ore under examination. The following operations may precede cyan idation.: 1. Amalgamation. Amalgam converted to bullion; tailing cyanided. 2. Coniuroi) or blanket strokes. Concentrate amalgamated; tailing cyanided. 3. Gmvity concentration, Concentrates may be a. Smelted. b. Amalgamated with or without cyanidation of tailing. c. Cyanided, raw or roasted; tailing cyanided. 4. Roasting. Calcine cyanided direct or after amalgamation or concentration on blanket or corduroy strakes. 5. Acid wash to remove soluble harmful constituents such as ferrous iron and copper compounds.

6. Aeration with lime solution. 7. Flotation. Concentration treated as in 3; tailing cyanided or discarded.

Laboratory testing of the foregoing operations may be conducted as follows: 1. Amalgamation. Several lots of the ore are prepared by grinding to such different sizes as may seem desirable. A 200-gram sample of ore is

28

CYANIDA'I'ION AND CONCENTRATION OF ORES

put into a bottle with 100 cc water, 1 gram NaOH, and 30 to 50 grams mercury. The bottles are placed on rollers and revolved for 1 to 2 hr. The pulp is then panned, and the mercury separated. Pouring the partially cleaned mercury from one clean beaker to another and washing with a strong stream of water will be found to facilitate final cleaning. A small amount of sodium amalgam will be found useful in causing all the mercury globules to coalesce. The tailings are dried and assayed. The mercury is dissolved in nitric acid, and any remaining gold washed, dried, and wrapped in lead foil with enough sil verfor parting and then cupeled. An alternative method is to dissolve the mercury down to a small globule, which is then run directly into the assay crucible. Sodium amalgam can be prepared by forcing small shavings of metallic sodium beneath the surface of mercury. Care should be taken to see that the mercury is dry. As the reaction is rather violent, due precautions should be taken in making the amalgam. Amalgamation tests can also be made in a grinding pan or an iron mortar, particularly when the ore contains much silver. Copper sulphate and salt may be added to the pulp. The ore may first be given a chloridizing roast. 2. Corduroy 01' Blanket Strokes. In testing the use of blankets or corduroy the pulp is wet ground to various sizes, and the pulp at a dilution of about 4 or 5 to 1 is passed over the strake set at an inclination of about 1~ to 1% in. per ft. Whenever possible, the strake should be not less than 10 to 12 ft. long; the width is not so important. If for any reason it is not possible to use such a length, or if the amount of ore is limited, some idea of the effect of the strakes can be obtained by repeateelly passing the pulp over a short strake. In laboratory tests the weight of concentrate is disproportionately large as compared with plant practice. The tailings are collected, weighed, and assayed. The blankets are washed, and the concentrate collected. To make a complete recovery of the concentrate, the blanket after being carefully washed should be burned and the ash added to the bulk of the concentrate. 3. Gravity Concentration. If enough ore is available, a laboratory Wilfley table may be used. For small lots of ore a gold pan or a small laboratory jig such as the Denver mineral jig is convenient and gives reasonably good results. Sufficient concentrate should be prepared so that the various methods of treatment as already noted can be tried. In hanclJing concentrate the experimenter should be particularly careful to avoid losing any fine free golcl. Both concentrate and tailing can be tested with and without regrinding. The concentrate should be dried at low temperatures to avoid oxidation


29

of sulphides, weighed, and assayed, and calculation made as to the ratio of concentration. "1. Roasting. In the laboratory, roasting experiments may be carried on by using fire-clay roasting dishes or heavy sheet-iron pans, preferably in an electrically heated muffle. The charges, period, and temperature should be carefully noted. Modern methods of temperature control have made possible greatly improved results in this field. If the ore contains much arsenic, the amount should be determined before and after roasting. If it is desired to reduce the arsenic to an extremely low amount, the calcine is reroasted with charcoal. The calcine may be passed over blanket or corduroy strakes and then cyanided after removing soluble acidic salts or neutralizing them with lime. 5. TVashes. Some ores contain soluble iron or copper compounds which cause a high consumption of cyanide. A water wash or a wash of dilute sulphuric acid or sulphurous acid may remove these compounds to permit of economical treatment of the ore by cyanidation. Washing has usually to be avoided following a chloridizing roast, however, owing to dissolution of gold chlorides, unless provision is made to recover them from the wash water. While acid pretreatment has been used in a few cases, most careful investigation should be made before adopting it. 6. Aeration. On some refractory ores and concentrates a preliminary intense aeration in a strong lime solution before the addition of cyanide has been found beneficial. The maximum solubility of lime, CaO, in the ordinary cyanide mill solution is about 2.5 lb. per ton of solution. However, as much as 50 to 100 lb. lime per ton of ore may be used, depending upon dilution and chemical consumption (see discussion of the Salsigne process in Chap. XIII). 7. Flotation. Flotation testing of ores containing precious metals may be for the purpose either of determining the applicability of the process to an ore under investigation or of determining the possibilities of improving the work of an operating flotation plant. In the latter case the work is best done at the mill itself, samples taken directly from various points' of the mill stream being used for comparison. In this way it is possible to obtain a pulp that has been ground under operating conditions as well as to use the mill water, which may be of great importance. When the tests are made to evaluate the application of the process to an ore, the work can, of course, be done anywhere. However, it should be made certain that the ore used for experimentation is thoroughly representative of that which is to be milled. Furthermore, if possible, a sample of the water to be used in the mill should be obtained, and its effects noted.

30

CYANIDATION AND CONCENTRA'l'ION OF ORES

Difficulties are frequently encountered owing to surface oxidation of sulphides, particularly pyrrhotite, and where flotation results cannot be duplicated aftm: an interval of several days 01' weeks, this condition may be suspected. On this account, it is also desirable to ship and store the ore sample at as coarse a size as is compatible with good sampling, crushing to under % to H in. only that portion which is to be used for immediate testing. One point that should always be kept in mind is the determination of the propel' place flotation may fill in securing the maximum net profit from the ore. On some ores an all-flotation treatment might be indicated; usually, however, better results may be obtained by combining flotation-if it be used at all-with other methods such as amalgamation, concentration, or cyanidation, Again, if all-flotation seems to be best, the possibility of separate treatments of different fractions of the pulp should be considered. For example, the so-called primary slime might advantageously be separated and floated in a separate circuit. Naturally, such points cannot be determined in advance, but if the operator will keep them in mind as the work progresses, he will be enabled to obtain a better perspective of the basic problem, which is to produce the maximum profit from the operation as a whole, and determine the proper balance between extraction and cost. The first testing should be confined to an all-flotation method so that the experimenter can become familiar with the ore. Careful observation, coupled with experience in testing, will enable the operator to interpret the various phenomena. It is desirable that he shall have had practical operating experience so that he may know the limitations of each step under consideration. Fortunately, it is generally true that, if the laboratory procedure follows sound operating practice, the flotation-mill results will equal or surpass those attained in the laboratory. The simplest flotation testing results when a single mineral or concentrate is desired. This is usually the case in copper ores when the copper occurs as one or more of the sulphide minerals. Such ores generally contain more or less pyrite. If the pyrite be present in such an amount that it would affect the grade of the concentrate, then differential methods of flotation are employed whereby the copper metals are caused to float and the pyrite is prevented from floating. Differential methods are also used in separating two or three minerals such as galena and sphalerite or galena, sphalerite, and pyrite. When making differential separations, it is important to know the degree with which one mineral may be intergrown or locked with another. This is determined by microscopic examination and chemical analysis. The distribution of the precious metals or their association with different base metals should be determined, as this has an important bearing upon the net value of the concentrate produced. For example,gold or silver associated with a lead concentrate yields a greater net return than an equivalent amount in a zinc concentrate. Within limits, the type of laboratory cell used is not a matter of great importance

EXAMINATION AND TESTING aFORE

31

as far as obtaining indicative results is concerned. That this is so is shown by the fact that different operators using different types of laboratory machirles with which each has become familiar will arrive at practically the same procedure. The important thing is to obtain a machine which is mechanically reliable and easy to clean and which permits quick work. The amount of material taken for a test will be governed by the amount" of ore available, size of machine, amount of floatable material in the ore, and various other factors. If the amount taken for a test is too small, any errors or effects of manipulation are unduly magnified. On the other hand, if the amount is too large, each test may consume too much time-not so much the actual flotation itself but the grinding or other preparation of the pulp and the drying and sampling of the products. Probably 500 to 1000 grams is the best all-round size of charge. If possible, however, several different sizes of machines should be available so that charges of 50 to 2500 grams can be used if desired, and more latitude is permitted for cleaning operation, especially where the ratio of concentration is high. Where differential separations are made or several cleaning operations seem called for, t.lie amount of pulp taken should be sufficiently large to allow the various operations to be made and to ensure resultant products large enough for assay. Broadly speaking, the reagents used in flotation may be divided into three classes. Various writers on flotation extend the division, and the nomenclature differs somewhat. However, the three following classes are generally recognized and accepted by flotation operators without confusion: 1. CONDITIONING OR "MODIFYING" AGENTS. These include such compounds as lime, soda ash, sodium silicate, sodium sulphide, sodium cyanide, sodium sulphate, sodium bichromate, copper sulphate, and zinc sulphate, which are in general use. There are others less frequently used. 2. PROMOTERS OR COLLECTORS. Under this heading are found the various xanthates, dithiophosphatos (aerofloat ), thiocarbanilid, fatty acids, and others classed as chemical promoters and also certain oils and tars. 3. FROTHERS. This class embraces such compounds as the various pine oils, cresols, alcohols, and other petroleum and coal- or wood-tar derivatives. It is not within the scope of this chapter to enter into a detailed discussion of the effects and use of the many reagents used in flotation. To those unfamiliar with the subject, .the suggestion is made that they obtain from the manufacturers or distributors of reagents data concerning their use. Then they should consult various books and publications in which details of flotation-plant practice are described. The U.S. Bureau of Mines, in cooperation with various mining companies, has issued a number of technical papers covering in great detail the flotation operations at most of the important milling plants in the United States. Publications of the various mining and metallurgical societies and technical journals contain many valuable articles. Manufacturers' catalogues also are well worth study. In Chap. XV considerable information will be found on the flotation practice at the various mills described. The amount of reagents used can be determined either by weight (dry reagent) or by volume (definite-strength solution), and the amount used is usually expressed in equivalent pounds per ton (2000 lb.) of original ore even though the operation involved is a cleaning step involving only a small percentage of the original feed. Lime is prepared as described under the sections on cyanide testing. Soda ash may be used either dry or in solution. If dry, it should be thoroughly dried to constant

32


weight und kept in [I stoppered bottle. The moisture content should be checked occusionully, [IS soda [Ish may absorb 15 to 18 per cent water and still appear quite dry. Wa.ter-soluhln salts such [IS zinc and copper sulphates and sodium cyanide should be used [IS solutions of definite strength-I, 5, or 10 per cent-and measured by pipettes. Xuuthn.tes [Ire best used as 1 per cent solutions and freshly prep[Ired each day [IS needed. Oils [Ire measured as drops from c[llil)mted pipettes or medicine droppers. A dropper is used for each oil and is calibrated l;y weighing 10 or more drops. It is convenient with some oils-e-purticularly steam-distilled pine oil-to have several droppers delivering different-size drops. This cnn be done by druwing out the tip of the dropper to [I onpillury and then breaking it off to obtain [I small orifice. In this way it is posslble to obtain a variution of 50 to 250 drops of pine oil per grum. The determination of the pH value of the water of the pulp should alwnys 1::e made for reference, and the effects of changing the pH noted. A good procedure is to take a sample of the dry crushed ore and grind it with distilled water in a porcelain jar mill with flint pebbles. The pulp is filtered, and the pH value of the filtrate determined. At the same time the water is examined in more or less detail for soluble salts. In order to obtain [In idea of the grade of concontrute [IS well [IS the recovery that may be obtainable in practice, the following tests may be made: Consider the simplest case where [I one-mineral concentrate is desired. After a few preliminary tests so that the operator has [I fairly comprehensive idea of such factors [IS degree of grinding, pH, reagents, and other pertinent de.tu, two tests are made. In the first one such conditioning reagents [IS may be required are added, and then three or four successive froths are removed and kept scparuto, staged additions of promoters and frothers being used. The several froths and the tailing [Ire dried, weighed, and assayed, and the distribution of the valuable mineral determined. All products should be examined under the microscope to obtain. an idea of the nature and amount of any locked or true middling grains. A oomparative test is then made under the same conditions, but only one froth is made, the same kinds and amounts of reagents being used [IS before. After the tailing has been diseharged from the cell and the cell cleaned, the froth is returned and refloated. Further addition of reagents to this cleaning operation may or may not be made ; only the judgment a.nd experience of the operator can determine. This operation of cleaning the froth may be repeated [IS often [IS deemed necessary. All products [Ire finally dried, weighed, assayed, and examined [IS before. The information so obtained may then be studied for the purpose of arriving [It an idea of a possible flow sheet, and the advnntages and disadvantages of the various types of circuits may be determined. However, [Ill such deductions should be made with caution and regarded only [IS indicative of whut to expect in plant practice. As flotation tests progress, the operator will obtain [In idea of the nature of the results being obtained by visual examination both of the froth and of the tailing. This is best done by vanning samples on [I white enameled vanning plaque. The plaque is then pI[Iced under [I microscope, and the product examined. The froth sample is readil y obtained by scraping it on to the plaque. The tailing or pulp sample may be drawn from the cell by means of [I gluss tube about ~:l: in. in diameter and 12 to 15 in. long, the ends of which have been closed sufficiently to hold the pulp after the operator has sucked it into the tube. Complete notes on every test [Ire essential. The operator should also record in detail observations such [IS the character of the froth, the effect of each addition of


33

reagents, and other ,phenomena. Nothing is too unimportant to be noted, so that months after a test has been made, the original experimenter or another can repeat the test and obtain substantially the same results; by reading the notes he can get an accurate and comprehensive picture of just what happened. . After the tests have been completed and a formal report is written, all pertinent data may be abstracted and assembled in a form designed to convey the necessary information, readily and accurately to the reader. With some ores the removal of the slime portion before flotation is extremely beneficial. Therefore, should an ore be encountered that seems to be difficult to float by usual methods, the effect of removing the slimy portion ahead of flotation should be investigated. Pructicully, slime may be considered as that portion of the pulp finer than about 50 microns .. In practice the slime is usually removed by means of a DOlT bowl classifier or a DOlT hydrosepurntor or by centrifugal means, such as the Bird centrifuge. In the laboratory the usual method is to agitate the pulp thoroughly and then siphon off the suspended portion at a rate corresponding to the settling rate of the largest particle desired in the overflow. This may be repeated if two-stage operation seems called for. Sometimes a "mud" or a "talc" froth may first be removed by flotation. Such froth will contain the greater part of the deleterious components with a negligible amount of valuable mineral. The tailing from this operation will then generally respond to usual methods of flotation. The removal of slime from a pulp often markedly improves differential separations, although the results may not economicully justify the added step. Caution should be used in recommending this procedure. After the laboratory work has revealed methods by which it is believed the ore may be successfully treated by flotation, the work should be reviewed in the light of the relationship of flotation to other phases of the entire milling process. There is, for instance, the question of the economics of direct oyanidation versus flotation and cyanidation of the concentrates only, and such studies involve many other phases of the treatment as ::I; whole. For example, there are grinding, classifying, concentration (tables or blankets and other gold-saving devices), thickening, and filtering. In the case of ores containing gold or silver it may be found that a combination of flotation and cyaniding may yield a greater net return than either one alone. The effect of the various reagents on classifying, .t.hickening, and filtering should be investigated, particularly if it is desired to recover the water for reuse. The equipment of a flotation laboratory may range from something extremely simple to quite an elaborate installation. The minimum requirements include some means of wet grinding and a flotation cell. More complete equipment may include various types of dry-grinding machines, wet-grinding units employing steel rods, steel balls and flint pebbles, several types and sizes of flotation cells, testing sieves, microscope, air compressor, and vacuum pump. The arrangement of the laboratory should be carefully considered so that there will be no lost motion in conducting a test. If possible, all wet work should be done on a table or bench covered with sheet zinc and sloping to a drain, so that any spills may be cleaned up readily. Coneuntption. of Flotation Reagents. The following information concerning reagents and quantities commonly used in the flotation of gold ores of no particular refractoriness is furnished by the American Cyanamid Company.

34


Pounds Addition agents: Soda ash Sodium silicate.................... Activating agents : Copper sulphn.te , . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Promoters: Aeroflout 15.......................................... Aeroflout 25.. . . . .. .. .. . . . . . . . . .. . . .. .. . . .. .. .. .. .. Aerofioat 31 Reagent 208 Reagent 301. Potassium amyl xantuate .. .' 'I'hiocarbanilid. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Frothers and froth stabilizers: Pine oil. , Cresylic acid. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Coal-tar creosotes , Pine-tar oil. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . ..

pCT

Ton

0.50 to 3.0 0.50 to 2.0

a.10

to 1.5

0.05 O.05 0.05 0.05 0.05 0.05 O.05

to to to to to to to

0.15 0.15 0.15 0.15 0.15 0.1.5 O.15

0.025 to 0.10 0.05 to 0.15 a.10 to 0.25 0.05 to 0.15

CYANIDE TESTS AGITATION METHOD

Procedure. The following procedure has been found satisfactory for carrying out cyanide tests by agitation. Winchester bottles of about 2.5- to 4-liter capacity are used. It is convenient to number the bottles and determine their tares, etching the figures on the bottle by means of hydrofluoric acid and then marking over the etching with a china- or glass-marking pencil. When a wet pulp such as the tailing from amalgamation or concentration tests is the material to be tested by cyaniding, the weight of the pulp is obtained by weighing bottle and contents and subtracting the tare. When the test is finished, the solids are dried and weighed, and the amount of solution used may be determined. Otherwise a weighed amount of ore is put into the bottle, and then a weighed amount of lime of known available lime content is added. Next is added a measured amount of a cyanide solution of known strength. The bottle with the pulp is then agitated. The agitating device (Fig. 4) consists essentially of a series of horizontal, rubber-covered, wooden rollers about 4H in. in diameter and 30 in. in length, and mounted on bearings with their axes parallel. They are spaced at about 7-in. centers and rotate at 50 to 75 r.p.m. When a bottle containing pulp is placed between two adjacent rollers, it is caused to revolve and the pulp is agitated. It is convenient to have a board set over the ends of the rolls and at a right angle to their axes with a notch midway between the centers of the rolls to engage the neck of the Winchester bottle.

EXAMINNI'ION AND TESTING

OF

ORE

35

The notches are at such a height that the bottles are tilted, allowing them to be filled with more pulp than if they rested on their sides and with no danger of the pulp's splashing out. Details of Testing. Enough ore for the maximum probable number of tests is prepared by dry grinding so that it all passes a certain. size, say 100 mesh. Samples for assay and screen analysis are taken. We shall assume that it is desired to determine the extraction obtained by treating a certain ore with a solution containing 0.05 per cent sodium cyanide, NaCN, and 0.025 per cent CaO (protective alkalinity), which means that the solution is to contain 1.0 lb. NaCN (free cyanide) and 0.5 oJ

FIG. 4. A device usee! in the testing laboratory of the DOlT Company for making bottle cyanide agitation tests. As many as 20 bottles can be rotated at one time.

lb. CaO (protective alkalinity) per ton of solution. The pulp is to be agitated at a dilution of 2.5 parts of solution to 1 part of ore for 24 hr. All crude ores will be found to vary in their lime-consuming power, and the extent of this factor should be determined prior to treatment. Provision must be made for the addition of enough lime not only to satisfy the lime-consuming power of the ore but also to produce the proper hydrogenion concentration during treatment. In determining the lime-consuming power of raw ores, the following procedure has been recommended by LaMotte, Kenny, and Reed in their book pH and Its Practical Application. Determining Lime Requiremenu: of Ores. Fifty-gram samples of the dry, ground are are placed in 8-oz. bottles with 200 cc distilled water. To the first sample bottle no lime is added, to the second the equivalent of 0.5 lb. calcium oxide per ton solution is added, and to the third, fourth, fifth, and

·36


other bottles 1.0, 1.5, 2.0, etc., lb. calcium oxide pel' ton solution is added. The bottles are stoppered and agitated for 1 hr., after which the pulp is filtered and the pH value of the filtrate determined. An analysis of the filtrate is then made to determine the amount of calcium oxide remaining in solution. When this value is known, the amount of lime consumed by the ore can be calculated. This procedure permits excellent control of the regulator, not only giving greatest economy but at the same time enhancing the efficacy of the other reagents used. A 400-gram sample of this ore is weighed out and put into a bottle. (Foul' hundred grams is a convenient amount, although any amount from 100 grams to 10 kilograms may be taken, bottles of suitable size being used.) The next step is to add the lime. It will be assumed that preliminary tests or other information indicate that the ore will require about 2.0 lb. CaO (100 per cent basis) pel' ton for neutralization during the first hour of agitation. This means that 400 grams will require 0.4 gram CaO. In addition to the lime for neutralization, enough lime will be needed to bring the solution to 0.025 pel' cent CaO. At a dilution of 2.5 to 1 there "rill be 1000 cc (or ml) of solution which, at 0.025 pel' cent CaO, "rill require 0.25 gram. The total lime to be added then is 0.40 + 0.25, or 0.65 gram. The hydrated lime has been tested and found to contain 69.0 pel' cent available CaO; therefore, we shall add 0.65/0.69 = 0.94 gram of hydrated lime. Water is added next, and the amount is determined as f 011 ows: 400 grams of ore requires 1000 cc of solution for a dilution of 2.5 to 1; 1000 cc of solution at 0.05 pel' cent NaCN contains 0.5 gram NaCN, equivalent to 50 cc of the stock 1.0 pel' cent NaCN solution. Therefore, add 950 cc water. The pulp is shaken in the bottle for a moment, and then 50 cc of the 1.0 pel' cent cyanide solution is added. Sometimes it is desirable to check the dilution, and this may be done readily by weighing bottle and contents if the tare of the bottle is known. The bottle is now placed on the rollers and agitated, the time being noted. At the end of a period varying from 7~ to 2 hr., depending upon the rapidity with which the cyanide or the lime is consumed, the solution is titrated for free cyanide and protective alkalinity. This is done by removing the bottle from the rollers and allowing the pulp to settle until some of the supernatant solution can be drawn off in a pipette. If the solution is clear enough, a 10- or 25-cc sample may be cautiously drawn off in a pipette and titrated direct. If the solution is turbid, about 50 or 60 cc may be drawn off, using a 100-cc pipette. The solution is filtered through a dry filter into a clean, dry beaker, and a sample is taken for titration (see Appendix B). Determination of Consumption of Cyanide and Lime. The following notes of a typical test will illustrate the method of calculating the consumption of cyanide and lime :

EX.A.MINATION AND 'TESTING OF ORE

37

Ore-400 grams. Lime added-0.94 gram hydrated lime = 0.65 gram CaO. Solution-1000 cc (050 cc H 20, 50 cc 1.0 per cent NaCN solution). Solution strength to be maintained at about 0.05 per cent N aC~ and 0.025 per cent CaO. 9:00 A.M.-start 10:00 A.M.-25 cc taken out for titration, giving 0.042 per cent NaCN, 0.022 per cent CaO. The solution is now brought back to a volume of 1000 cc having a strength of 0.05 per cent N aCN and 0.025 per cent CaO. ()75 cc at 0.042 per cent N aCN = 0.41 gram N aCN; 1000 cc of solution should contain 0.50 gram N aCN; therefore, add 0.09 gram N aCN or 9 cc 1 per cent N aCN solution and 16 cc H 20. J 975 cc at 0.022 per cent CaO = 0.21 gram CaO; 1000 cc of solution should contain 0.25 gram; therefore, add 0.04 gram 100 per cent CaO or 0.04/0.60 = 0.058 gram hydrated lime. 1:00 p.M.-25 cc taken out for titration, giving 0.048 per cent NaCN and 0.02 per cent CaO. The cyanide .strength is close enough to that desired so that no correction is necessary. The protective alkalinity, however, is lower than desired and is raised as follows: 075 cc at 0.02 per cent CaO = 0.195 gram CaO. This amount subtracted from 0.250 gram (1000 cc at 0.025 per cent) = 0.055 gram; therefore, 0.055/0.69 or 0.08 gram hydrated lime is required. As the pulp may be expected to continue to consume lime for the next few hours, a slight excess over the theoretical amount may be added, say a total of 0.10 gram. Then, 25 cc of water is added to bring the volume of solution back to 1000 cc. 9:00 p.M.-25 cc taken out for titration, giving 0.045 per cent NaCN and 0.023 per cent CaO. To bring the solution to the desired strength there are required 0.50 - 975 X 0.00045 = 0.06 gram NaCN and 0.25 - '()75 X 0.00023 = 0.026 gram CaO. As the pulp is now to be left to agitate overnight without further attention, add slightly more than these amounts to take care of consumption, say 0.08 gram N aCN or 8 cc 1.0 per cent solution and 0.05 gram hydrated lime. Also added 25 minus 8 or 17 cc water to bring the solution volunle to 1000 cc. 9:00 A.M.-finish of 24-h1'. agitation. The solution titrates 0.049 per cent NaCN and 0.023 per cent CaO. Results may be calculated as follows:

NaCN Ou; NaCN In cc 1% solution 25 cc at 0.042% NaCN 0.01050 gram NaCN 25 cc at 0.048% NaCN = 0.01200 gram NaCN cc 1% solution cc 1% solution 25 cc at 0.045% NaCN = 0.01125 gram NaCN cc 1% solution 1000 cc at 0.049% NaCN = 0.40000 gram NaCN 0.52375 gram NaCN or 0.67 gram NaCN 0.67 - 0.524 = 0.146 gram NaCN consumed by 400 grams ore. This is equivalent to 0.146 X 5 = 0.73 lb. 100 per cent NaCN per ton ore. 50.0 9.0 8.0 67.0

38


CaD Tn

CaD Out

25 cc at 0.022% CaD = 0.00550 gram 0.940 gram hydrated lime 25 cc at 0.020% CaD = 0.00500 gram 0.058 gram hydruted lime 25 cc at 0.023% CaD = 0.00575 gram 0.100 gram hydrated lime 0.050 gram hydrated Iime 1000 cc at 0.023% CaD = 0.23000 gram 0.24625 1. 1'18 or 1.148 X 0.69 = 0.792 gram CaD o 792 - 0.246 = 0.546 = gram CaD consumed by 400 grams are. This is to 0.546 X 5 = 2.73 lb. of 100 per cent CaD per ton of are.

Cl10 CaD CaD CaD

equivalent

It may not be necessary or particularly informing to carry out tests to such a detailed degree as that shown. Each experimenter may determine for himself the extent to which it is desirable to correct solution strength and dilution during testing. Rather than maintain the solution strength at approximately a predetermined figure, some experimenters prefer the more simple method of not making corrections of the solution during the period of agitation, starting with a solution high enough in cyanide and lime so that at the end of the period the solution is considered to be adequate. Such a procedure, however, is open to objection on the ground that the solution may be too strong at the beginning and too weak at the end, and these conditions may lead to erroneous deductions. Series Testing. After a preliminary test has indicated the consumption of cyanide and lime, several series of tests may be outlined in which the effect of changing one variable is determined. These variables may be cyanide strength, protective alkalinity, fineness of grinding, time of agitation, dilution, addition agents (lead of mercury salts, bromocyanide and others). Each series should embrace 4 to 6 deg. of the variable or as many as may be believed to be desirable. Finally, several tests may be made using various combinations of the best conditions as determined by the series tests. Addition Agents. With certain ores it may be found that the addition of a small amount of a lead or mercury compound is beneficial, either in purifying the solutions or in increasing the extraction or both. Lead in the form of litharge, as the nitrate or as the acetate, and mercury as the metal or oxide may be used. Mercury is now seldom used. . . It is always worth while to make parallel tests with and without one of these agents. One-half to 1 gram of litharge or 4 or 5 grams of mercury added to 400 to 1000 grams of ore is ample. In practice, probably less than 74 lb. of litharge per ton of ore is generally enough. As the effect of litharge may vary greatly with the amount used, it is of great importance to determine accurately the critical amount to use.

39

EXAMINA'.rION AND TESTING OF ORE

Sometimes it is desired to determine the effect of bromocyanide on certain gold ores, particularly those containing tellurides. As bromocyanide is an unstable compound, it must be freshly prepared before use, and as it and its vapors are extremely poisonous, due care should be used in its preparation and use. When bromocyanide is being used, the protective alkalinity should be kept at the lowest possible point. In this connection a parallel test of the same low alkalinity but without the bromocyanide should be made to make sure whether possible improved results should be ascribed to bromocyanide or to low alkalinity. The reason for maintaining a low alkalinity when bromocyanide is used is because BrCN is rapidly decomposed by alkali. The reaction is usually illustrated thus: BrCN

+ 2KOH

= KBr

+ KCNO + H

20

In his 111tmual of Cyanidation, E. M. Hamilton says this: The usual method of making the reagent for laboratory use is to add a strong solution of cyanide to bromine (and not conversely) until the brown color is just discharged: KON

+ Br2 =

KBr

+ BrON

The quantity of BrON may be determined in a working cyanide solution by acidifying with hydrochloric acid, adding an excess of potassium iodide and titrating the liberated iodine with decinormal sodium thiosulphate. J. E. Clennell in his Chemisiru of Cyanide Soluiions, 2d ed., gives the following reactions: BrON + HOI = HON + OlBr OlBr + 2K1 = KOI + KBr + 12 1 cc O.IN thiosulphate = 0.00529 gram BrON The BrON is added to the cyanide solution to be used for extraction purposes in the proportion of about 1 BrON by weight to 4 of KON.

Change of Solution. It' is often of interest to determine the effect of a change of solution upon the extraction (rate and amount) and the chemical consumption. The procedure of effecting such a change will depend primarily upon the dilution of pulp during agitation and the dilution to which the pulp will settle after standing a short time. It is preferable to remove the solution by decantation rather than by filtration. Precise manipulation is necessary in effecting a change of solution. Decluctions as to the effect of this procedure should be made with caution, the practicability of reproducing the effect in plant operation being kept in mind. Grinding in Solution. The small samples available for laboratory experiments cannot be expected to reproduce exactly conditions that will exist in large-scale milling practice. Actual size reduction of the ore can be dupli-

40


cated, of course, but closed-circuit grinding with a high circulating load would be very difficult to reproduce on a laboratory scale. Particularly is this true when grinding in cyanide solutions is to be practiced. This is not to say that grinding in cyanide solutions should not be performed in the laboratory. On the contrary, such grinding is likely to add considerably to the knowledge accumulated in testing, and it should be done. But the data thus gathered should be weighed with mental reservations and regarded as subject to some variation. As a matter of fact, laboratory grinding in solution is standard practice at some testing plants; and the information gathered is regarded as valuable. Perhaps the nearest approach to solution grinding in the laboratory is to grind in water, classify to desired sizes, and settle. Then cyanide solution and lime are added to make up a solution of the desired strength. This pulp may be then agitated for a time, ~~ to 1 hr., to check the approximate dissolution in the mill-classification circuit. After this the pulp may be thickened to the dilution at which the agitation and aeration are to take place. Cyclic Use oj Solution: After preliminary batch tests have indicated the optimum conditions such as fineness of grind, dilution, strength of solution, and time of agitation, tests should be made in which cyclic operations of . agitation, thickening, decantation, washing, and precipitation are carried out. The important point here is the reuse of the solution. It may be found that fouling of the solution takes place and its efficacy diminishes to a greater or lesser degree. If such be the case, it is essential that some method of reactivating or purifying the solution be determined to avoid the necessity of discarding too large an amount. Testing for Reprecipitation. Reprecipitation of the dissolved gold may under certain circumstances take place in the ore pulp before the solution is separated from the solids. The most frequent cause is the presence of carbonaceous material in the ore, but it is not generally realized that pyrite and pyrrhotite in the presence of low free cyanide concentration can also act as effective precipitants for gold. A suitable test is to agitate 500 grams of the ore with a pregnant solution containing a low cyanide concentration and the same amount of gold as contained in 500 grams of the ore. It was found, for instance, in the case of one ore tested by this method that 85 per cent of the gold was precipitated from the pregnant solution by the ore. PERCOLATION METHOD

In cases where fine grinding is not necessary to liberate the gold values, it is often possible to classify the coarsely ground ore into a sand and slime product (at, say, about 200 mesh) treat the former by so-called percolation


41

leaching and the latter by the agitation method that has been described above. For laboratory tests a convenient apparatus for percolation leaching consists of a 3-ft. length of Ui in. inside diameter lucite or glass tube held by clamps in a vertical position on the laboratory bench. The upper end is left open, while the lower end is provided with a rubber stopper fitted with a ~:4:-in. glass outlet tube, to which is connected a length of rubber hose. Several shallow drainage slots radiating from the center are cut on the face of the. stopper, a circular piece of light cotton filter cloth slightly larger in diameter than the stopper is placed over the slots, and the stopper inserted into the tube. The filter cloth will wedge between the stopper and sides of the tube and be held firmly in place. The end of the outlet tube, which is provided with a regulating clamp and small glass nozzle, is arranged to carry the effluent solution into a Winchester quart bottle. Solution is supplied to the tube by means of an inverted stoppered bottle conveniently placed on a shelf above the bench and having a large outlet hose the end of which is submerged below the solution level in the tube. As the level drops, air is admitted to the bottle and enough solution flows out to submerge the tube and reestablish the level. In carrying out a test, the tube is filled to a depth of about 24 in. with the deslimed sand, and the feed bottle filled with about 2 liters cyanide and lime solution of known strength. The outlet clamp is closed, and the solution run into the percolation tube to a level about 6 in. above the surface of the ore. The end of the inlet hose is fixed at about this level. The outlet clamp is then adjusted to give a flow of 5 to 10 cc per min. into the effluent collecting bottle. Under the constant head, and if the slime content of the sand is low,' there is little difficulty in maintaining this flow, while the supply is automatically maintained, as above described. When all the solution has run through (at the end of 4 to 8 hr., depending on feed rate), the charge should be allowed to aerate for some time, while the effluent solution is titrated for cyanide and lime content and made up to strength by suitable reagent addition. This solution is then transferred to the feed bottle, and the cycle of operations repeated. The leaching operations should be continued for several days at least, at the end of which time one or two water washes are applied and the sands discharged, dried, and assayed. In connection with laboratory percolation tests, E. M. Hamilton" remarks as follows: One of the principal difficulties in such percolation tests lies in the fact that the charge is so much shallower than a working charge that there is not sufficient fi

Mtuiual of Cyanidation, p. 226, 1920.

42

CYANIDNl'ION AND CONCENTRATION OF ORES

head to overcome the Ci1pilbrity of the interstices, so thnt, even with sand coarse enough to percolate very rupidly, the level of solution will not f'all much below the level of the sand. The result of this is t.hut the charge is not uerutad (us it is in pmctice), by the air following the solution down into the interstices, between each wash. This mav be overcome by applying a vacuum under the filtering medium nfter the solution has ceased to percolate by gravity. In this Wi1Y the residual solution is drawn off and the air follows it down. The charge should then stand for several hours before the next wash is applied, This procedure is more imp ortunt t hnn it may seem, since a difference in cxtruct.ion of 20 to 30 per cent hus been in some instances observed according to whether the vucuum was applied or not.

PRECIPITATION

Method of Gold Precipitation. The following method of precipitation of gold in pregnant cyanide solutions has been adopted by the Ore Dressing and Metallurgical Laboratories, Bureau of Mines, Ottawa. 6 Pinch-cock

"D

JJ

--To sccto»

FIG. 5. Laborutory gold-precipitation apparatus. Canada.)

iBureoai of Mines,

Ottawa,

The apparatus consists of three Erlenmeyer flasks of 3000-cc capacity (see Fig. 5). Flask A is for evacuation of ail' from the pregnant solution, flask B is for precipitation, and flask C is for the barren solution. These flasks are connected in series by tubing and are airtight. The "barrensolution" flask C is connected to the suction. The filtering apparatus for gold precipitate is prepared as follows: The Buchner filter is connected to a suction, and three sheets of filter paper are placed on the perforated bottom of the filtering funnel. The filter paper is covered with a layer of asbestos. The latter is packed down and then covered with a layer of about 716 in. of diatomaceous silica. A small piece of glass is placed on top of the layers below the opening in the rubber stopper ; this prevents the solution entering the filtering apparatus from washing out a depression in the asbestos and diatomaceous silica layers. 6

C.M.J., July, 1939.


43

In the "evacuation" flask A is placed 330 cc of clarified pregnant solution. The flask is shaken by agitation apparatus in such a manner as' to give the solution a centrifugal motion. This type of motion spreads the solution along the walls of the flask, thus exposing a larger surface area and also a thin layer of solution to the vacuum. 'When the air has been evacuated from the solution (this point is readily observed by the disappearance of froth caused by the agitation), the flask is turned over and the solution passed into the "precipitation" flask B, in which has ~een placed a known amount of zinc dust, and agitated for 15 min. At the end of that time, the rubber tubing between flasks i1 and B is clamped by means of a pinchcock E and the connection broken between the two flasks. The glass tube D connecting flasks Band C is pushed down to the bottom of the flask, and a small amount of air is admitted into the flask B by means of the pinchcock E. This lowers the vacuum in flask B, and the solution passes through the filter into flask C. This method is not comparable to the Merrill-Crewe method used in mill practice. Hence it is necessary to run comparative tests against ores which are known to give very low barren solutions in mill practice. SETTLING AND FILTRATION TESTS

Settling-test Procedure. The principles of settling-test procedure and the application of formulas to determine sizes of continuous thickeners are relatively simple and are briefly described below. However, detailed manipulations under specific problems and the application of test results to successful milling practice call for experienced interpretation. 7 The following formula expresses the relationship between the settling rates of pulp at various dilutions, in terms of thickener area required: A = I

1.333(F - D) R X sp. gr.

where A = square feet per ton of dry solids per 24 hr. R = settling rate in feet pel' hour of a feed with F dilution sp. gr. = specific gravity of liquid F = weight ratio of liquid to solids for the rate R D = weight ratio of liquids to solids in discharge The zone requiring the greatest unit area is found by applying this formula to pulp of different densities, ranging in dilution from feed to discharge density. It is this zone which determines the area that must be provided for the pulp being tested. It is important that settling tests to determine the size of equipment for a cyanide circuit be carried out in cyanide and 7 For a theoretical discussion of this subject see "Thickening-Art or Science?" by E. J. Roberts, T.P. 2541B, New York meeting of il.I.M.E., February, 1948.

44


lime solution, following the actual period of agitation desired and using the required amounts of cyanide and lime. Thickening Capacity. The following formula is used to determine the volume provided in a tank in the thickening zone. Such volume depends directly upon the period of detention required for the sludge to reach the desired density:

v=

4T(G - sp. gr.) 3G(8 - sp. gr.)

where TT = volume in cubic feet required for thickening pel' ton of solids pel' 24 hr. 8 = average specific gravity of thickened pulp during compression period sp. gr. = average specific gravity of solution G = average specific gravity of solids in pulp T = period of detention in hours Use of Reagents. Cyanide pulps are usually flocculated to some degree owing to the maintenance of normal protective alkalinity, and this results in line settling and a fairly clear overflow. In some cases, however, additional lime or other flocculating reagent must be added, and among these, caustic starch is one of the most effective. In summarizing the results of tests using this reagent the authors" of an A.I.M.E. technical paper state: Starch solutions capable of flocculating finely divided solids suspended in water can be prepared either by heating under pressure in the range of 100 to 160°C. or by causticizing starch paste. Maximum efficiency with a noncaustic solution is attained when the reagent is prepared at 140 to 145°C. Causticizing temperature depends on the strength of caustic solution used. At 25°C. an efficient reagent can be produced with a 2.5 per cent solution of commercial NaOH. Starch reagent can be prepared most economically by causticizing or heating a 5 per cent starch paste with thorough mixing and diluting. Any starch can be used to prepare a flocculating reagent, but potato starch is recommended. Solutions prepared by heat alone will retain their properties for '3 days. Causticized solutions retain their properties for 2 weeks or longer. The use of these flocculating reagents appears to have a wide range of application.

Filter Tests and Calculation. Table 4 illustrates the technique of laboratory filter tests and the deductions gained therefrom. The conditions were an ore pulp of 42 pel' 'cent moisture content, filtered on a leaf of 0.5 sq. ft. capacity at room temperature. The objective was a reasonably dry cake with maximum capacity and a filter to handle 100 tons of dry solids per day. 8 "Flocculation and Clarification of Slimes with Organic Flocculants," T.P. 1052, A.I.M.E., February, 1939.

45

EXAMINATION AND TESTING OF ORE TABLE

4.

FrvrER-TEsT DATA

Test Conditions

I

1

Cloth used ............................ Feed volume, cc ...............' ....... 1000 Cake forming, min .................... 2~
H 1115 21 880 6% 828

I

2

I

3

Twill 26· 500 650 1 % 25 25 1 7~ 23 20 Slightly Slightly "cracked cracked

%

~
Satisfactory 560 725 23 22 423 566 2 3 1370 1200

4

400 ~i

25 ~i

25 Good ~
458 23 357

IH 1520

The results of tests 2 and 3 were selected, and, a 40 per cent safety factor being allowed," a filter area of 220 sq. ft. for 100 tons per day was recommended. "The factor of safety to apply to any individual test is a matter of judgment based on experience and will vary considerably.

CHAPTER III

Coarse Crushing JJ1 ethotls for the coarse crushing of ore range [rom. a simple combination of a grizzly and one crushe: to an elaborate arrangement in series of grizzly, jaw crusher, screens, and rolls 01' cone crushers. Crushing is done underground at a few large mines. In general, the jaw crusher is the most suitable primary machine, yet many gyratories do the first breaking. There is a wide choice of these machines. Wearing parts are of special steel, especially hard-faced at the point of greatest abrasion. For secondary and tertiary crushing, gyratories ancZ cone crushers are used, the latter particularls; for final reduction, but this can also be done by rolls, which give a uniform product. Interposed in a large crushing plant should be grizzlies, screens, feeders for crushers, magnets for picking tip loose iron and steel, and deoicesjor weighing the ore. The problem of ore comminution involves equipment that will handle the large pieces of rock resulting from mining operations and tum out a product of the relatively small particle size necessary for successful gold extraction. This size reduction is almost always carried out in two successive steps; coarse crushing (dry) followed by fine grinding (wet). Dry crushing is seldom carried below about H-in. particle size, experience having shown that the breaking of material beyond this point is more satisfactorily accomplished by wet methods. On the other hand, there is a trend today toward feeding considerably coarser material to rod mills which function to some extent as wet rolls. An efficient crushing plant almost invariably employs the principle of stage reduction, with grizzlies, trommels, shaking or vibrating screens interspersed between stages to separate out the material that is already fine enough to by-pass. the next step. The grizzly-stationary, rotary, or reciprocating-is suitable for the first stage, but thereafter the trommel or screen is more positive in its separation. Few trommels are used, except those on dredges and in sand and gravel plants. SELECTION AND POSITION OF PRIMARY CRUSHERS

Coarse crushing or breaking is the primary mechanical step in ore reduction. It is frequently done underground in large plants such as the 46

COARSE CRUSHING

47

McIntyre and Hollinger mines. Where sorting is done, it is usually at the shaft mouth, although central plants may be used. . As regards primary crushing machines, the choice is limited to two types-the jaw crusher and the gyratory. . The jaw crusher is, of course, a reciprocating-action machine, crushing rock by direct pressure between a fixed plate and a swinging jaw. The gyratory crusher, on the other hand, consists essentially of a circular shell with inner sides inclining toward a central orifice. A central shaft passes vertically through this opening and is hung centrally from a spider, which spans the opening at the top. This shaft is eccentrically moved at the bottom and is equipped with a conical crushing head, which operates between the inclined sides of the shell, crushing rock between the head and

FIG. 6. Principal parts of a Buchanan type B jaw crusher.

shell. The crushing operation in this machine is continuous and rapid. It is able to clear itself when completely buried in its feed. In selecting the proper type of primary crusher a balance must be established on the basis of the largest size of rock to be crushed and the total quantity of material to be handled in a unit of time. It is also to be borne in mind that the product of the gyratory is much more uniform in size than that of the jaw crusher. Cone crushers are a modification of the gyratory type in that the conical crushing head is eccentrically mounted on a vertical-drive shaft that is supported entirely from below. The top crushing mantle is spring loaded to take care of tramp iron and overloads, and the clearance between the crushing faces is adjustable by rotation of the top assembly which is attached by a heavy threaded connection to the frame of the machine. The Nordberg Manufacturing Co. makes the Symons standard cone crusher for

48


normal-size reduction and the Symons short-head crusher for fine crushing operations. On the Rand the older plants operate jaw crushers for crushing ahead of stamp mills, but in the newer ones, jaw crushers break the ore prior to sorting, and gyratories are used for the finer reduction ahead of tube or ball mills. Cone crushers are used in a few plants, and probably more will be installed. While jaw and gyratory crushers compete for primary crushing, gyratory and cone crushers share the secondary field.

FIG. 7. Sectional view of a Symons short-head crusher.

TRAMP STEEL

Magnets are used for the collection of tramp steel mainly at large mills. At small plants most of the large pieces of steel are picked out of the 'ore by crusher attendants. At the McIntyre Porcupine, for example, a large magnet at the Symons crusher supplied by a direct-current generator uses 24 to 30 amp. at 250 to 300 volts, depending upon the temperature. A heterogeneous collection of iron and steel results. Permanent magnets of the new high-permeability Alnico alloys are being used to an increasing extent for iron removal in a number of industries, and it is probable that this practice may eventually be adopted in metallurgical work.

COARSE CRUSHING

49

FEEDING CRUSHERS

There are roughly four periods in the normal cycle of coarse crushing where feeders are not employed: (1) The breaker receives a large charge of ore, (2) operates under choke feed with the power at a maximum, (3) gradually discharges the crushed ore, and (4) runs empty. "This is not .the most economical method yet is the practice in many crushing. stations. Some method of feeding a crusher regularly is preferable. This may be done by means of a feeder or gate to the bin above the machine or by a heavy chain, a pan feeder, a shaking grizzly, or an apron feeder, which occupy little head space. The objective is to ensure a regular feed to keep the crusher steadily occupied and thus flatten the power curve. Figure 8

FIG. 8. Apron-feed conveyor for Superior jaw crusher.

shows an apron feeder-jaw crusher arrangement. Cone crushers and rolls, particularly, may be fed by belts or drum feeders; cone crushers sometimes are fed from chutes. A typical Ross chain feeder consists of six lengths weighing 57~ tons. The feeder is driven by a 5-hp. motor through a bronze worm reducer, with a sprocket chain-gear and pinion arrangement. This gives a speed of 19 r.p.rn. and a capacity of 175 tons an hour. CRUSHING PRACTICE FIRST STAGE

1. Crushing Underground. This practice has been adopted on a number of properties, particularly in Canada. At the Kerr-Addison and Hollinger properties the· ore is crushed underground in jaw crushers to 4- to G-in. pieces before being hoisted to the surface. The same scheme

50


is followed at the Pamour-Porcupine mines and at the McIntyre Porcupine, where the underground crushers make about a minus 7-in. product. 2. Crushing on Surface. At the Golden Cycle custom plant, Colorado Springs, treating Cripple Creek and other ores, a 5J~-ft. Symons cone crusher is fed with ore of 3H to 4 in. maximum size which has passed over a vibrating screen having 0.18- by O.g-in. openings. The Symons produces 100 tons an hour of J~-in. material for further comminution. At the Premier mill, British Columbia, the mine-run ore is dumped on a grizzly of 80-lb. rails, spaced at 12 in. A Stephens-Adamson apron feeder delivers the grizzly oversize to an 18- by 3D-in. jaw crusher set at 3 in. At the Ross shaft of the Homestake mine, the primary crushing is done by two Allis-Chalmers size 8 gyratories, They are driven by 70-hp. motors and reduce the ore to 4H in.

FIG. 9. Crushing action of rolls, showing 4: 1 reduction ratio. SECOND STAGE

Secondary crushing may be done by jaw or gyratory breakers or by cone crushers. In the surface plant at the McIntyre-Porcupine a 7-ft. Symons cone crusher (see Fig. 11) set at H6 in., crushes material from the mine breaker. A 200-hp. motor with V-belt connection drives the crusher. Second-stage crushing at Loreto, Pachuca, Mexico, is as follows: 4-in. ore from the first crusher. combined with the undersize of a grizzly above it, is passed over two SPnons 4- by 8-ft. rod deck screens, tl!.~ oversize of which is fed to two Symons 5H-ft. standard secondary crushers, making a %-in. product. Each is driven by a 225-hp. motor and V belt and consumes 172 hp. At the Premier mine, British Columbia, 3-in. ore from the primary breaker passes over a Niagara screen with 'l-in. openings. The oversize is fed direct to a size 6 McCully crusher set at 1 in. At the Ross shaft of the Homestake mine, the 47~-in. ore from the pri-

COARSE CRUSHING

51

mary breakers passes over a grizzly to two 7-ft. Symons cone crushers, set at 1~4 in. in open circuit. ' THIRD STAGE

The third step in crushing, or tertiary reduction, depends upon what is required. Two-inch or smaller pieces from the secondary crushers may be fed direct to stamps, rod mills, or ball mills, but it is more usual to use a third dry crusher to reduce the ore to H- or %-in. size before wet crushing, particularly in the case of ball milling. For such work the Symons shorthead machine is well suited, just as the standard machine is for coarser red~ti~ J

CANADIAN PRACTICE

Hollinger. The Hollinger crushing plant is one of the largest and most up-to-date plants in Canada. Of steel and concrete construction throughout, high headroom is provided so that, with the exception of a single lift through shuttle conveyors to Nos. 2 and 3 screens, there is gravity flow to all machines (see Fig. 10). The ore is crushed underground to minus 6 in. and hoisted to 140-ton surface bins. From there it is fed to a 7-ft. Symons cone crusher after passing over a double-deck screen that removes 17~-in. undersize from the crusher feed. The product of this crusher plus the product of all other crushing units in the plant is then conveyed to the secondary screens which are situated at the top of the building. These 6- by 14-ft. vibrating screens separate out the 47:2- by ~16-in. undersize for storage in a 200-ton bin, from which point it is fed by roll feeders into six 5- by lO-ft. tertiary screens that make a final undersize (H6 by H in.) for transfer to the mill bin. The oversize from the secondary screens is crushed in a 57:2-ft. Symons shorthead crusher set at % in., and the oversize from the tertiary screens crushed in a set of 78- by 20-in. rolls set at H6 in. Both machines are closed-circuited with the secondary screen, and in this respect the flow sheet differs from the four-stage crushing at McIntyre Porcupine, where each crusher is separately closed-circuited with its respective screen. The crushing plant has a maximum capacity of 440 tons per hour and carries a circulating load of about 2: lover the final screens and 1200 to 1400 tons per hour over the secondary screens. The plant is completely equipped for dust control with some 40,000 cu. ft. per min. of air drawn through the cyclones, and 60,000 cu. ft. per min. of air through roof vents. On account of the large volume of air being displaced, it is quite impossible to attempt to heat this building in the winter, and there is only a difference of a few degrees between inside and outside temperatures. Heated cubicles are provided for the operators.

52


McIntyre. The crushing plant at the McIntyre Porcupine has been changed over in recent years from a three-stage to a four-stage operation. The mine ore is first crushed underground to about 7-in. size, then hoisted Mine are l2ass sy-sfem

t

Vibrating feeder

I' ~--/ 'I 48 x 60 jaw crusher I underground/

t

-

Vibrating feeder

t t Vibratingfeeder t

Crushed are. 120cket

Measuri(lgand loading hOlll2erS

lJ

4 -6Yz ton skins

-.~

140-ton 13urface stora9.e bin

t

-

15' x 8' Oquble-deck vibrating screen 3" and 1f4~q:.2l2enin9.. ,

I"

(Oversize from both decks)

l,T

I

t

11I Sy-mons fane crusher ~-----.... 2 - 6 ' X

f

ee)

(1:t4 x 1:t4 UnderSize

1 " T~" 14' VIbrating . screens f 4 u& • x '16 0l2emng

,--r-

(Oversize)

(Undersize)

,

t

2-80 ton surqe bins

200 ton screenedarebin

t -

t

8elt feeders

Roll feeders

,f

U~-5'y"mons sharI-head crushers

~ ~~

t

6 - 5'x 10' Vibration screens ~61/ x ~ 1/ ol2eninq

+

--,-

(Oversize)

(Undersize)

+

t

Surqe bin

To mill arebin

-r-

8elt feeder

t

18 1/ x 20 1/ Rolls set to 3:-;6 " 125 r.12.m. FIG. 10. Flow sheet of crushing plant. Ontario, Caruula.)

I"~ conveyors's no not sown h )

f,/8,e I

(Hollinger Consolidated Gold Mines, Ltd.,

53

COARSE CRUSHING

to the surface and stored in a 700-ton ore bin. From here it is fed to a standard 7-ft, Symons cone crusher which discharges to two Symons rod deck screens with the ~~-in. slots set transversely to the flow of ore. These screens are closed-circuited with a 5~-ft. Symons short-head crusher, and the undersize is passed to a bank of six 4- by 6-ft. Hummer screens having H6- by %-in. openings. The oversize passes to the fourth stage of crushing-Traylor 78- by 18-in. rolls, the product of which is returned to the screens, while the undersize passes to the mill bins. See Table S for the analysis of Crusher house products. Mine OPe

t

Trax.lor jaw crusher 36"x 48" (underground)

--

f

-

Mine shaft (6-ton skli2s)

f

,

-

100-ton are bin

t

1 Standard Sx.mons cone crusher

t

2-Sx.mons rod deck vibrating screens 1"slots

.r;

(Unddrsize)

I

5fE' Short -head simons cone crusher I

r

,

3

II

=t

(Oversize!

.. l

u

5

II

•

--r

...-----'>- 6-4 x6 Hummer screens 16 x

:.-'8

openm{L

(Undersize)

!!Px!!K.fJLo rolls 18 x 18

n

tI

To mill bins FIG. 11. Flow sheet of crushing plant. Canada.)

(McIntyre Porcupine Mines, Lid., Ontario,

RECENT DEVELOPMENTS

Hadsel Mill. The Hadsel mill (see Fig. 12) is a new type of wetcrushing device which has found limited but interesting application. A large wheel, up to 24 ft. in diameter by 4~ ft. wide, is fitted with internal buckets. As the wheel revolves, the buckets lift the rock to the top of the mill where it is dropped onto stationary breaker plates. The ore is thus crushed by its own impact on the plates. Ore as large as 12 in. is fed to the mill. The following notes are quoted from the article' 'Three Years of Operation with a Hadsel Mill" by George A. Bell. E. and lJ!I.J., Vol. 141, No.1, p. 32,

·54

CYANIDATION AND CONCENTRATION OF ORES TABLE

5.

SIEVE ANALYSIS OF CRUSHER HOUSE PRODUCTS AT MCINTYRE

7-ft. standard Symons" Mesh Tj·ler Standard Feed, Per Cent

+ -

-

-

-

-

-

-

6 4 2 1.5 1.050 0.742 0.525 0.371 3 4 6 8 10 14 20 23

..," - vi) - 48

- 65 -100 -150 -200

+ + + + + + + + + + + + + + + + + + + + +

6 4 2 1.5 1.050 0.742 0.525 0.371 3 4 6 8 10 14 20 28 35 48 65 100 150 200

In. In. in. In. in. in. in. in. mesh mesh mesh mesh mesh mesh mesh mesh mesh mesh mesh mesh mesh mesh mesh

22.4 17.1 26.1 5.8 6.2 4.1 2.6 2.4 1.6 1.2 1.2 0.8 0.6 0.6 0.6 0.7 0.6 0.5 0.5 0.4 0.4 0.4 3.2 -100.0

Discharge. Per Cent --

0.8 1.7 10.2 16.5 14.8 11.0 8.5 5.6 4.6 3.6) 2.4 2.0 1.8 1.9 1.6 1.3 1.3 1.1 0.9 0.9 7.5 -100.0

57~-ft.

short-head Symons Feed, Per Cent

Discharge, Per Cent

78- by 18- in. Traylor rolls

Feed, Per Cent

Discharge, Per Cent

Hummer vibrating screens

Feed, Per Cent

-- -- ---- --

Discharge, Per Cent

--

1.7 4.0 19.1 44.9 25.7 1.9 0.1 0.1 0.1

0.1 0.8 0.1 0.6 1.5 11. 9 1.6 31.0 2.2 6.5 8.8 2a.0 14.8 6.7 10.9 8.5 18.3 12.7 12.0} 0.7 4.3 23.9 22.0 17.2 3.2 19.6 20.9 16.6 11.6 9.0 8.0 17.2 2.2 4.6 0.1 5.1 4.6 12.3 1.3 1.5 3.2 8.8 1.0 3.4 0.7 7.8 0.1 2.5 2.7 0.9 0.5 6.7 0.1 0.8 2.4 2.4 0.4 5.2 0.1 1.7 1.7 0.3 0.6 4.0 0.1 0.5 1.3 1.3 0.2 3.3 0.1 1.0 1.3 0.5 0.3 2.9 0.2 1.0 1.1 0.3 0.4 1.0 2.2 0.1 0.3 0.8 0.3 2.0 0.1 0.7 0.8 0.3 0.3 1.4 6.0 7.1 15.3 2.5 3.5 -----------100.0 100.0 100.0 100.0 100.0 100.0

* The

7-ft. Symons discharge is separated on Symons screens into +7~ in. and in. fractions. These screens are in closed circuit with the 57~-ft. short-head Symons crusher. Symons screen undersize is Hummer screen feed. Hummer screens are in closed circuit with the rolls. -7~

which describes an installation at the Demonstration Gold Mines, Ltd., Bagnio, Philippine Islands: The ores are composed of andesite, quartz, breccia, calcite, and kaolinized material which forms a sticky gangue that can be put through an ordinary crushing plant only with great difficulty. . The capacity of the mill operating on a 12-in. feed and crushing through a l"in. punched plate was 20 tons per hour, and power consumption about 3 kw.-hr. per ton of are. The manganese-steel breaker plates lasted approximately 4 months, and. other wearing parts 12 months.

55

COARSE CRUSHING

In conclusion, the author states: 1. The mill is ideal for crushing wet and sticky ores in one stage. 2. The mill is mechanically sound and will crush to any desired size of ball-mill feed efficiently. 3. Operating cost per unit of actual work is less than with standard machines." 4. There is consideruble saving in equipment.

oJ

FIG. 12. Principle of operation of the Hardinge-Hudsel mill.

\

CHAPTER IV

Sorting and Sanzpling When ore in mining becomes diluted by country rock because of narrow siopes because of weak wall rock, 01' when it naturally contains barren material which is distinguishable from the ore, sorting and rejecting the worthless rock may be economical. This is done in a casual manner at some small mines and on a lm'ge scale at others, as was the case at Alaslco-Fumeas: (40 per cent refected), 01' covering whole districts, as at Kolar, India (10 per cent), and on the Rand (9 per cent). Sorting at a gold OJ' silver mine may mean its existence, as at the Alaska-Juneau, 01' it may resul; in lower costs because less ore is crushed and treated to recover the same amount of gold, as at Cripple Creek, at Kolar, and on the Rand. 01'

SORTING IN THE FLOW SHEET

The place at which sorting will be done is mainly dependent. upon the size of the mine-run ore. If the ore is not in too big chunks, the sorting belt may be placed below the grizzly or trommel to receive the oversize after it has been sprayed with water. The grizzly undersize is transported direct to the storage bin. If mine-run ore is in too large pieces, it should be broken to 4- to 7-in. size, sprayed, and then fed to the sorting belt. The wash water containing the fine material sometimes carries enough gold to be worth saving; therefore provision should be made for sampling and assaying it also for its propel' disposal. EXAMPLES OF SORTING

Sorting on the Rand. The sorting of waste (low-grade) rock has been accepted practice for many years on the Witwatersrand. Low labor cost and the peculiar structure of the "banket" ore are both contributing factors. In some plants tube-mill pebbles as well as reject rock are removed from the belts and delivered to separate bins, while other plants merely pass the rock after sorting to grizzlies or trommel screens, where 6-in. oversize is separated out as grinding media and the undersize passes to the secondary crushers. In his presidential address before the Chemical, Metallurgical and Mining Society of South Africa, A. Clemes' states: There has been little change in the last 10 to 15 years in the established method of removing waste rock from slow-moving conveyor belts feeding the primary (and 1

"Modern Metallurgical Practice on the Witwatersrand."

56

SORTING AND SAMPLING

57

possibly secondary) fine crushers. On most plants ore passing through jaw breakers set at, say, 5 in. and retained on grizzlies (or screens) set at between 2 and 33:1: in. is subjected to waste sorting. Where intensive sorting is deemed necessary (say 15 per cent of the ore delivered), it is usual to arrange for two sets of sorting belts, one carrying minus 5 in. plus 3 in. material and the other minus 3 in. plus 2 in. The admixture of these sizes is, obviously, detrimental to intensive sorting, but where sorting of waste rock is not important, the capital cost of installing additional screening and belt conveyor equipment may be uncalled for. "Time studies" of the efficiencies of native sorters have been carried out on most mines with some measure of individual improvement, but efforts in this direction are largely nullified by the fact that often only poor quality or transient labor is allocated to this work, whilst the sortability of ore varies widely. In general, it may be accepted that under normal conditions n~tives sorting coarse waste can average about 10 tons per 8-hr. shift, falling to well below 5 tons per shift under conditions of intensive sorting of the smaller sizes. Purely from the angle presented to the metallurgist, benefication of the ore to the mill still seems profitable. Assuming a total cost of metallurgical treatment of 2s 9d per ton with a residue value Is 9d (0.20 dwt.), it appears worth while to discard waste rock valued at 3s (0.35 dwt.) at a cost for sorting and dumping of about Is 2d per ton; equivalent to a gain of 4d per ton. Such a simple arithmetical justification of waste sorting is not, however, completely acceptable today. Native labor is in short supply, and a number of reduction plants are working below capacity; possibly, more mines could now profitably take in waste rock yielding, say, 80 per cent extraction of 0.35 dwt. or 2s 4d per ton. Thus, under present and forecast conditions of shortage of man power, one is inclined to wonder if the continuance of waste sorting on some producing mines can be justified, also if the capital cost of providing for waste sorting-conveyor belts, waste bins, waste-disposal equipment, etc.-should be incurred in new plants.

Sorting at Randfontein. The ore is first screened on seven 8- by 3-ft. Tyrock double-deck screens, with 3-in. round hole openings on the top deck and l,72"-in. square mesh screen on the bottom deck. The undersize of these screens goes directly to the mill bins, and the oversize to washing and sorting. The +3-in. oversize of the top deck, which ranges in size up to 14 in., is washed and sorted on five 36-in. wide by 1I8-ft. long belts, from which waste and primary tube-mill pebbles are sorted. The oversize of the lower deck passes to two similar belts from which waste and secondary tubemill pebbles are sorted after washing. The washing is done by sprays on the lower end of the sorting belts, using 525 gal. per min., and drainage from the belts or washing fines are dewatered in two simplex Don classifiers with rake product going to the mill bins and overflow 84 per cent minus 200 thickened in two intermittent settling tanks, from which the thickened pulp is pumped to the secondary grinding circuit classifiers. The sorting belts nm at 25 ft. per min., and on an average 40 native boys pick about 60 tons per hour 'of waste and tube-mill pebbles. In 1946 the

58


waste amounted to 2.05 per cent of the crude ore and had an average value of 0.152 dwt. per ton. The cost was 13.06d (21.8 cents) per ton of waste. Sorting at McKenzie Red Lake, Canada. The crushing plant will handle the mill tonnage (225 tons per day) in 8 hr. The ore from the 200-ton mine ore bin is fed by a swing-hammer feed regulator to a 30-in. link-belt feeder delivering to a 2-in. bar grizzly at an angle of 35 deg. The fines drop to an 18-in. conveyor belt, and the oversize to a 36-in. sorting belt, where it is water washed by sprays prior to sorting, which is done under fluorescent light. The average production per day in 1945 was 209 tons, of which 35 tons was sorted to waste. The total cost per ton of ore mined and milled was $7.656. THE ECONOMICS OF SORTING

The question as to whether in any given instance sorting is justified is strictly a matter of economics. The cost of installation and operation of a sorting plant and the inevitable loss of some gold value, however small, in the rock discarded must be balanced against the saving in milling cost resulting from the elimination of low-grade rock. In this connection, R. D. Lord in "Milling at Preston East Dome," C.J\!I.J., August, 1941, gives the following formula: . Tons milled = A Tons sorted out = B Tons mined = A + B Cost of mining per ton mined = C dollars Cost of milling per ton milled = D dollars Cost of sorting per ton sorted = E dollars Value of waste as backfill = F dollars Grade mined Grade sorted G ra d em!'11 e d

= = =

m. p

. meA

+AB)

-

pB

Recovery from mined are = x (as a decimal fraction) Recovery possible from sorted low grade = y (as a decimal fraction) Recovery from milled ore 2

=

xmi~ + ;; - Y~B ni

+

-p

(as a decimal fraction)

2 To obtain the above formula we have that the gold recoverable from the total are hoisted is mx(A + B) and that the gold recoverable from the material sorted out is ypB. Then for the ore reaching the mill the gold recoverable is xm(il + B) - ypB out of the total gold reaching the mill which is meA + B) - pB. This makes the fraction recovered expressible as

mx(A

meA

+ B) + B)

-

ypB

- pB

59


Without sorting: Operating cost = A(C + D) Recovery = xmA dollars A(C + D) xmA

Cost in dollars per dollar in gold recovered With sorting: Operating cost of sorting

=

Recovery

=

=

=

xm

BE (in dollars )

. I = . ki Cost 0 f pic cing per ton nnner

Cost of milling Tutulcost = C(A

+D

C

A BE +B

AD

+ D) + BE + AD + (ypB xm(A + B) - ypB meA + B) meA + B) _ pB X A xm(A + B) - ypB

BD) p!3

X A

Cost in dollars per dollar gold recovered C(A

+ B) + BE + AD + (ypB

- BD) xm(A + B) - ypB Then the maximum value of muterial that can be discarded without increusing the cost per ounce of production occurs when C

+D

= CA

xm

+ CB + BE + AD - (ypB xmA + xmB - ypB

- BD)

This equation reduces to p=

xm(2D - E) y(C

+ D + xm)

xn[2D - ( E - F)] yCC

+ D + xm)

when introducing F, the value of the waste t1S backfill, Example: Making x = 0.974m = $11.40 D = $0.592 E = $0.490y = 0.800 C = $3.14 F = $0.30 0.974 X 11.4(1.184 - 0.190) p = = $0.93 = 0.024 oz. 0.8(3.14 0.592 0.974 X 11.4)

+

+

SORTING BY SINK-FLOAT

As yet not applied to the sorting of gold ores, this method is extensively used for the elimination of low-grade material in the treatment of coal and various "metallic" and "nonmetallic" ores." Based simply upon the principle of "floating" the waste in a heavy media suspension having a density intermediate between that of the desired "sink" and rejected "float," the method presupposes that an appreciable difference exists (at least 0.2) between the specific gravity of the ore constituents. Ordinarily as applied today, the method will handle minus 2-in. plus ~{6-in. feed (occasionally as 3 The sink-flo~t plant of the Bunker Hill & Sullivan Mining & Smelting Co., which is milling a silver-lead ore at Kellogg, Idaho, is a cnse in point.

60


fine as 10 mesh), but improved methods are under development which may soon make it economically possible to handle material as fine as 100 mesh.

,

f .

-.-

No. 2 unit

No.lum!

24 ''x 36 "A.C. jaw crusher.

----r-

15"x36" Universal jaw crusher

+

--r-

Magnetic l2ullev -f-~

Mag(letic l2uIlex.

4 'x 6' Vibrating screen I

4'x6' Vibrating screen I (Ove/size)

, I 4 Standard Symons crusher

f

42" Vezincutter I

Y5

f

I

Cut 30 "x 14" rolls set to 30" Vezincutter

f

ve"

•

4P.;Cut

I

>-5 Cut 20 "x 12"rolls setto 4mesh Jones riffle

t

I

15lb. sample

~

mer

1 Bucking room Final !P Ib sample . crushed to pass 150 mesh I

,.

Sample held in partable bins until settlement is made I I

l'

~4 Cut

~ Cut 30"x 14" ralls set to~" 30" Vezincutter

~---------


61

The principal reason why gold ores have not thus far been handled by this method lies no doubt in the fact that it is rare to find coarsely mineralized gold ores in which the values are highly concentrated in a heavy fraction after crushing to the size range above indicated. For details as to the techniques employed in sink-float processes, see Taggart's Handbook of 111ineml Dressing, Secs. 11-104; The American Cyanamid Company, Ore Dressing Notes Numbers 12, 13, and 14.; and "Heavy Density Separation-a Review of Its Literature." Quart. Colo. S. ofM., Vol. 45, No.1, January, 1948. SAMPLING MILLS

Various methods are employed for obtaining a representative head sample of the ore passing through a treatment plant. Some of these are described in Chap. XII (see also Taggart, "Handbook of MineralDressing," Secs. 19-54). For the sampling of large tonnages, however, and especially where custom ore is being milled, special equipment is provided for the purpose. The method used involves crushing the ore to a certain size which depends upon the size of cut to be taken, taking a relatively large cut with a Vezin-type sampler," crushing this cut to a smaller size, recutting, and so on, until a few hundred pounds of final sample is obtained, all the rejects being stored, usually in a separate bin, until such time as it is convenient to run the lot through the mill. An example of such a plant is that of the Golden Cycle Corporation of Colorado Springs, which hanclles mine and dump ore from the Cripple Creek district of Colorado. A flow sheet showing the steps in the sampling procedure is given in Fig. 13. ., Taggart, Handboolc of Mineral Dressing, Sees. 19-30, Wiley, 1945.

CHAPTER V

Fine Grinding Fine grinding is the last step in the reduction of an ore prior to concentration (gravity 01' flotation) 01' cyanidation. Practice varies, depending upon the type of ore and the amount of reduction required. In addition, some of the older properties continue with methods that perhaps are not considered the best in light of recent improvements but that cannot be economically changed because of capital outlay. Present grinding practice is closely linked with classification, so that some overlapping of subject matter occurs 'in Chaps. V and V I. In this chapter some of the theory of grinding, d(fferent types of equipment, and flow sheets are discussed. Concentrate regrinding is discussed in Chap.

IX. EQUIPMENT

Most of the tonnage milled today is ground in one of the following types of equipment or a combination of two or more: ball mills, tube mills, rod mills, and stamps. Chilean mills and Huntington mills are used only in a few isolated cases today. BALL AND TUBE MILLS

All types of mills that consist essentially of rotating cylinders with flat or conical ends with balls, pebbles, or rods used as grinding media arc included here. Ball Mills. The term "ball mill" is generally used to refer to a cylindrical mill whose length is less than, equal to, but not much greater than its diameter. It was initially developed for relatively coarse grinding, but by using it in closed circuit with a classifier its use has been extended for fine grinding. Ball mills have shells of cast iron or steeel plates and are carried on hollow trunnions. Ore is fed through a scoop, drum-type, or combination feeder at one end and is discharged from the opposite trunnion. Ball mills may be arbitrarily classified into two types, according to the method of pulp discharge. In high-level or overflow mills the pulp level builds up until it overflows and discharges through the trunnion. Highlevel discharge mills are made by a large number of manufacturers throughout the world. Low-level mills are typified by the Allis-Chalmers and 62

63

FINE GRINDING

Marcy (see Figs. 14 and 15) grate-discharge mills. The discharge end is fitted with grates; between the grates and the end of the mill are radial lifters which act as a pump to lift the discharge to the hollow trunnion. Drive is by spur or herringbone gear, direct connected 01' belt driven.

<\>F.::::~=~~;;;:;=======~g-' ~ ~

1i ~ f-J-::;;;-~+4J.....~-~~~~~~-~..;J (l,

~~~~~~~~~l@J. The Oorr Duplex /

sid. model C classifier

Note.

/~~::'f:·-:r

Pullt?ys can be furnished in this posifion if desired

FIG. 14. Allis-Chalmers ball granulator in closed circuit with

DOlT

duplex classifier.

FIG. 15. Sectional view of Marcy ball mill in operation.

64


Ball mills are built in sizes ranging from small laboratory mills to a present maximum of 12 ft. diameter by 12 ft. long, the latter requiring close to 1000 hp. Liners are usually of manganese steel, of chrome steel, or white iron, 3 to 6 in. thick. Corrugated and shiplap construction is commonly used to increase the grinding action. Conical Mills. The Hardinge mill (see Fig. 16) differs from most ball mills in that conical ends are added to the cylindrical portion of the mill. The cone at the feed end has a larger open angle than that at the

FIG. 16. Hardinge conical mill fitted with discharge grate.

'opposite end. Its makers state that the large balls concentrate neal' the feed end of the mill where the coarsest ore collects and the smaller balls act on the finer ore. Rod Mills. Rod mills (see Fig. 17) follow the general dimensions of tube mills with diameters from 3 to 6 ft. and lengths from two to three times their diameter. They differ from ball mills in that steel rods 301' 4 in. shorter than the mill length inside the liners are used as grinding media. Rod mills are often run on tires and rollers instead of trunnions or on one trunnion and one tire and set of rollers. Low-level discharge is obtained on Marcy rod mills by having a beveled annular ring at the discharge end. A stationary steel door fits close to this

65

FINE GRINDING

beyeled ring and serves to hold the rods in the mill while pulp discharges ' between the mill and the door. Tube Mills. The distinction between tube mills and ball mills is not so marked as their names indicate. Mills from 4 to 6 ft. in' diameter and from 16 to 22 ft. long are usually termed tube mills. This was the first type of rotary mill for metallurgical purposes. Because of the necessity of completing the grind during one passage (open circuit) of the ore through the mill, it was built with a large length-diameter ratio. The tube mill is still largely used in South Africa and to some extent in North America for fine grinding generally following some other primary mills. oJ

FIG. 17. Sectional view of Marcy rod mill in operation.

Tube mills are usually supported on hollow trunnions, the feed entering through a feed scoop at one end and discharging through the other. Drive is by a large gear fitted over the mill shell. Various types of liners are used, as in ball mills. Flint pebbles and hard waste rock were first used as grinding media, but present practice is to use small steel balls (less than 2 in.) or, as in South Africa, a mixed steel ball and rock load. GRINDING MEDIA

Steel balls ranging from % to 5 in. in diameter are used. Rods range from 1~ to 4 in. in diameter and should be 3 to 4 in. shorter than the inside mill length. Tube mills are usually fed balls smaller than 2 in., whereas

60


4- or 5-in. balls are more commonly used for ball-mill grinding. A much higher grinding capacity is obtained in tube mills by using steel media instead of pebbles, but in making such a conversion serious consideration must be given to the ability of the steel shell to withstand the greater loading. Approximate ball loads can be estimated by assuming 300 lb. per cu. ft. of ball volume and a total load equivalent to 40 to 45 per cent of the mill volume. Rod loads average about 40 per cent of mill volume, and a figure of 400 to 425 lb. per cu. ft. of rod volume should be taken. A comprehensive discussion of the theory and practice of wet grinding is to be found in Taggart's H anclbook of il1ineral Dressing, Sec. 5. The following excerpts from this work are of interest. Experience indicates that rods are superior to balls for feeds in the range from to 1 in. maximum when the mill is not called upon to finish at sizes finer than 14 mesh. Balls are superior at coarser feed sizes or for finishing I-in. feeds to 28 mesh of grind or finer because the mill can be run cataracting and the large lumps broken by hammering. In an operating mill a seasoned charge, containing media of all sizes from that of the renewal or replacement size down to that which discharges automatically, normally produces better grinding than a new charge. It is inferred from this that a charge should be rationed to the mill feed, i.e., that it should contain media of sizes best suited to each of the particle sizes to be ground. . .. Usual practice is, however, to charge a new mill with a range of sizes, based on an assumed seasoned load; thereupon to make periodic renewals, at various sizes dependent upon the character of the circulating load, until optimum grinding is obtained; and thereafter to make required renewals at the optimum size. A coarse feed requires larger (grinding) media than a finer feed. . .. 'The smaller the mesh of grind the smaller the optimum diameter of the medium. This relationship is attributed to the fact that fine product is produced most effectively by rubbing, whence maximum capacity to fine sizes is attained by maximum rubbing surface, i.e., with small balls. A practical limitation is imposed by the tendency for balls that are too small to "float" out of the mill and by the high percentage of rejects when renewals are too small. The usual materials for balls are chilled cast iron and forged steel, for rods, highcarbon steel, (0.8 to 1.0 per cent carbon) all more or less alloyed. . .. Mild steel rods are unsuitable for the reason that they bend and kink after wearing down to a certain minimum diameter and snarl up the whole rod load. The hardened steel rods break up when they wear down and are removed at about 1 in. or left in an eventually discharge in small pieces. ~~-

LINERS

All rotary mills must be fitted with some kind of replaceable liners. Chrome steel, manganese steel, and white iron are generally used. Shapes designed to give a corrugated or shiplap surface to the interor of the mill are often used to prevent slippage of the ball load.' Pocket liners are 1 This factor is thought to be the major cause of liner wear and loss of grinding efficiency. See "Ball Mill Liners" by W. L. Howes, Trans. A.I.M.E., 153, 1943.

FINE GRINDING

67

also common. These liners have pockets III which the balls become lodged to form the wearing surface. Rubber liners have been tried experimentally but have not been adopted by the industry. According to Taggart," no fully satisfactory method of holding the liners in place was worked out, utility was limited to "fine feeds and small balls, mill capacity was reduced, and while a slightly higher grinding efficiency was shown in comparative tests with steel liners, there was no indication that possible increased wear for rubber would. offset its far greater cost. Silex liners with flint pebbles for grinding media are sometimes used where iron contamination must be avoided. ., THEORY AND PRACTICE BALL AND TUBE MILLS

Grinding Action. The grinding that takes place in mills of this type i3 usually ascribed to two actions, impact and attrition, although some authors do not believe that a sharp line of demarcation can be drawn between the two actions. In rod mills there is line contact between the rods, there is less grinding by impact, and the action resembles that of crushing rolls. As a result, a rod-mill product usually contains a greater percentage near the limiting size with less extreme fines than ball or tube mills. Grinding-mill Capacities. It is generally accepted that in the case of a given are the amount of grinding which can be accomplished through a particular mesh by .a ball or rod mill is dependent upon two factors: 1. The size of the grinding media. 2. The net" power input to the mill. In selecting the correct grinding media it is important that the rods or balls supplied be large enough to break the largest particles of are in the feed, and as already discussed, a seasoned load composed of balls of all sizes, which is the condition found in a mill that has been operating for some time, gives better grinding efficiency than a new charge. Assuming that the correct grinding media has been established, the net power input can be varied by varying 1. The total weight of the grinding media. 2. The speed of the mill. 3. The size of the discharge opening. 4. The percentage solids in the pulp. Taggart, Handbook of M'ineral Dressing, Sees. 5-25, Wiley, 1945. refers to the power the motor draws less the energy lost in gear, bearing friction, and ail' resistance. Grinding studies at the Lake Shore mill confirm the findings of other investigators that horsepower input varies as the 2.6 power of the mill diameter, other conditions being constant. 2

a This

68


The volume of the charge is limited to a maximum of about 50 per cent of the mill volume. If the charge is too large, its center of gravity shifts too neal' the axis of the mill and the power input falls. The speed of the mill is limited by what is known as the "critical speed." This is the speed at which (assuming no slippage) the charge starts to cling to the liners, or to "centrifuge." It is given by the formula. 7G.50 N=

Vd

where N = revolutions per minute d = diameter in feet inside the liners In practice, the speeds used vary from GO to 80 per cent of critical, depending upon the individual preference of manufacturers and operators. The size of the discharge opening governs the amount of pulp held inthe mill, and too large a pulp volume reduces the power input for the reason discussed above. The percentage of solids in the pulp is usually maintained at GO to 75 pel' cent, the principle being to keep the volume percentage of solids as high as possible without loss of mobility of the charge. The correct proportion of water present will depend on the kind of ore being handled, slimy ores in general requiring a higher dilution than ores that have a low slime content. The size of mill required for a specific grinding problem will depend on the character and size of the feed and the product desired and whether open01' closed-circuit grinding is desired. An accurate estimate of capacity can be made only by an engineer familiar with the proper evaluation of the factors involved. For rough estimating purposes Table G gives approximate capacities grinding to 48 and 100 mesh for several size mills. Connected horsepower is also shown. These figures are for what would normally be considered average siliceous ore and for nominal circulating loads of 2 01' 3 to 1. These capacities may be reduced by as much as 50 per cent in the case of a hard, tough ore which is highly resistant to grinding, and for this reason considerable thought has in recent years been given to methods for determining the relative grindability of different ores and to correlating laboratory figures with plant performance. F. C. Bond has published comprehensive grindability data (T.P. 2180, A.I.M.E; July, 1947) based on work carried out by the Allis-Chalmers Manufacturing Co. and grindability tests are a regular part of the testing procedure. of the Dorr Company at the Westport, Conn., laboratories. Open- and Closed-circuit Grinding. When the tube mill was first introduced, grinding was done in open circuit; i.e., the ore was ground to

69

FINE GRINDING

pass the limiting screen size by one passage through the mill. It was found, however) that if sufficient time of contact between the' ore and grinding media were provided to ensure that no unground particles (or oversize) discharged from the mill, an excessive amount of· fines were produced. This meant that the ore was ground much finer than necessary and mill capacity was correspondingly reduced. . TABLE

6.

:MILL CAPACITIES AND POWER REQUIREMENTS'"

48 mesh grind, tons per 24 hr.

100 mesh grind, tons per 24 hr.

Size of millj

Type

OJ

%-in. feed

~2-in.

~1-in.

feed

feed

Approx,

Connected horsepower H-in. feed

grinding media charge, tons

---Cylindrical ball mills

Conical mills

ball

4 5 7 8 9 472ft. 7 ft. 8 ft. 10 ft.

by by by by by by by by by

3 ft. 6 ft. 6 ft. 6 ft. 8 ft. 16 in. 36 in. 48 in. 66 in.

Size of mill

Rod mills

4 6 7 8

ft. ft. ft. ft.

by by by by

25 100 320 500 1000 34 225 400 900

30 120 360 600 1200

10 35 125 200 350

20 72 200 300 536 23 150 275 625

20 mesh grind from f-in. feed, tons per 24 hr.

48 mesh grind from l-in. feed, tons per 24 hr.

200 720 1200 1700

120 410 720 1000

8 ft. 12 ft. 15 ft. 16 ft.

* From manufacturers' catalogues. t Diameter times length. (In the

20 to 60 to 150 to 200 to 375 to 15 to 100 to 175 to 400 to

30 75 175 225 400 20 125 200 450

Consumed horsepower

38 to 130 to 175 to 335 to

50 150 270 350

2.0 6.5 12.0 17.5 32.0 2.0 11.5 19.0 35.0 Approx. grinding media charge, tons

6.5 22 33 55

case of the conical mill this refers to the

cylindrical section.)

The difficulty was overcome by placing a classifier in the circuit to separate out oversize from the mill discharge and return it to the mill feed. In closed-circuit grinding no attempt is made to finish the grind in one passage through the mill, but every effort is made to remove finished material as soon as it is released, thus reducing overgrinding and preventing the fines from hindering the grinding action on yet unreduced particles. In this way the tonnage that a given mill will grind is much greater than it is possible to grind in open circuit. By using wide classifiers with high raking capacity, circulating-load

70


ratios are now being carried to 4: 1 or higher. The direct result of the increased capacity is reduced power, liner, and grinding media consumption per ton of finished ore. There is, of course, a limit as to how large a circulating load can be carried in practice. While capacity continues apparently to improve, though at a decreased rate, it becomes increasingly difficult to move the growing volume of material through the system. The different types of classifiers used are discussed in the chapter on "Classification." There is some controversy in the literature as to the definition of ratio of circulating load. The term used by most millmen is the ratio of sand tonnage returned to the mill to the tons of original feed. If the mill-classifier circuit is fed into the classifier instead of into the mill, the sand contains oversize from the original feed as well as oversize from mill discharge, and thus the definition is not entirely accurate. The ratio of circulating load can be calculated from screen analyses by using the following formulas: 1. When the raw feed enters the mill directly:

· .1 • IoacI ratio . C irculatingwhere d = o = s = 2. When

d- - d0 s-

cumulative percentage on any mesh in the mill clischarge cumulative percentage on same mesh in the classifier overflow cumulative percentage on same mesh in the classifier sand the raw feed enters the classifier: Circulating-load ratio

where

=

= .f -

s-

dO

f = cumulative percentage on any mesh in the raw feed

0, are the same as in 1 The usual procedure in using these formulas is to calculate the circulating-load ratio for several size screens and discard any that appear out of line, averaging the remainder.

d, s,

FLOW SHEETS

There are many types of flow sheets in use today. The tendency in new mills is to crush relatively fine (% to H in.). Single-stage ball mills in closed circuit with classifiers are used for grinds coarser than 48 mesh, but when a finer product is desired, two stages of ball mills in closed circuit with classifiers is usual. Efficiency must necessarily be sacrificed to . some extent in small mills by capital requirements, and even greater reduction ratios are justified in a single-stage grinding unit.

71

FINE GRINDING TABLE

7.

RAND PRACTICE: GRINDING :MILL DATA,

Grinding loud, tons

Mill shell dimensions, ft.

A C E B F G H D G H A B C E D F G H J

16 20 16 8 10 9 9 20 20 20 16 16

20 22 20 16

20 20 15

24.5 26.0 21.0 20.0 17.5 21.5 21.5 26.0 25.5 25.5 19.5 24.5 26.0 30.0 20.0 20.5 25.5 28.0

350 250 340 202 400 145 106 240 160 195 280 308 250 181 240 215 165 164 245

Prim. Prim.

Prim. Prim. Prim. Prim Prim. Prim. Prim.

Prim. Sec. Sec. Sec. Sec. Sec. Sec. Sec. Sec. Con ct.

7 ,j

12 25 34

14 14 Nil Nil Nil 4

... t 6 6

Nil Nil

Nil Nil

New feed

Product, per cent

12 90095.8 2.9 62.5 10 480 1.0 53.0 198.0 17 140072.5 7.2 42.7 o 950 4.5 .... o llOO 5.0 .. "!. o 300 5.8 .. o 260 4.2 . 15 400 4.0 4-1.0 14 270 . . 5.3 .... 14 240 ... 4.2 .. 13 450194.0 4.6 .. t 540 . 30.0 .... 162.526.9 10 48053.031.0 5.0 12 62542.739.415.0 15 40044.038.0 2.0 19 550 35.0 .. 15 540 49.4 .. 15 530 49.1 . Nil 110 6.4 144.7 0.0

1948*

Production,

Power,

tons per

lcw.chr. per ton

mill duy

26.9 300 31.0 234 17.9 452 30.5 35.0 49.8 .. 51.0 .... 38.0219 49.4 51.0 72.5 303 79.6 .... 70.0 238 61.2 192 75.0 182 63.6 73.7 72.0 99.6 60.0t

216 156 291 248 330 130 122 149 129 128 215 290 193 150 158 170 153 131 80.0§

7.020.9 29.0 9.419.1 28.8 4.413.5 21.0 5.7 21.9 6.5 21.7 8.8 20.2 11.4 24.4 10.719.1 28.8 10.6 22.2 14.5 27.3 11.116.6 23.3 10.5 ... 19.0 9.418.8 23.2 5.216.9 21.6 10.723.6 27.2 7.0 22.7 5.5 19.4 5.6 22.5 39.9 73.1t,54.8§

• Production of pebble and composite mills includes pebble consumption. t Composite load: tonnage not given.

meSh} FI t . . dimg 92 per cent §t 200 350 mesh j a ut.ion coneen t rate grm

.

IDlllUS

350 mes h .

FIG. 18. Grinding section at McIntyre Porcupine; note flotation cells mills and classifiers for removing free gold.

bet~een

tube

72


With the large classifiers used for high circulating loads it is quite often necessary to use some kind of auxiliary device to complete the closed circuit. A large motor-driven scoop lifting the mill discharge to the classifier has been successful. Rod mills are used to some extent for relatively coarse grinding, and the trend today is to use them in open circuit for the primary grinding stage.' In North America stage grinding is often carried out with ball mills as primary grinders and tube mills for fine grinding in the second stage. OLDER GRINDING EQUIPMENT

Stamps. Although the stamp mill could be classified as a fine crusher, it is included in this chapter, as its usual duty corresponds approximately to that of a primary ball mill. Stamp mills were built to parallel the operation of a mortar and pestle, working continuously and on a large scale. Ore is fed into a mortar and is crushed by the dropping of the stamp on a die at the bottom of the mortar. The crushed ore discharges through a screen in the side of the mortar. The shoe that forms the wearing surface on the dropping stamp is attached to a steel stem and is replaceable. The stem is lifted by a cam operating against a tappet which is bolted to the stem. A common camshaft activates usually five stamps in a battery. The most highly developed mill of this class is called the "California stamp mill," a complete description of which can be found in Richard and Locke's Text BODle of Ore Dressing," p. 38, 1940. As a matter of historical interest the following account of Colombian practice as late as 1935 describes the primitive type of stamp mill from which the modern machine was developed. Milling was done in unique, crude wooden stamp mills developed by the ingenious Antioquenan miner. Made entirely of hand-hewn hardwood (except for cast-iron shoes, several bolts, and a few nails) these molinos Antioquenos have a stamp duty of approximately 0.4 tons per 24 hr. They are powered by overshot water wheels, 18 to 24 ft. in diameter, mounted directly on the 18- to 24-in. wooden camshaft of the mill. Up to 56 drops per minute can be obtained with a water-wheel speed of 14 r.p.m, The stamps, 6.H by 77~ in. by 14 ft. in dimensions, weigh 450 to 500 lb. including the cast-iron shoe. The mills are usually built with three stamps to the mortar box and as many as three sets (nine stamps) per mill. Battery-box screens are usually made of tin from 5-gal. gasoline cans perforated with a small nail. Stamp guides, cams, and the hardwood camshaft bearings are lubricated with beef tallow.

Modern stamp mills employ stamps weighing from 850 to 1500 lb. in the United States and 1500 to 2200 lb. (the Nissen stamp) in South Africa. <1 See "A Study of Present Day Grinding" by L. E. Djingheuzian presented at the Montreal meeting of C. I M. and M., April, 1949, where the paper "Fine Crushing with a Ro'd Mill at Tennessee Copper Company" by J. F. Meyers and F. M. Lewis is reviewed.

FINE GRINDING

73

Stamps drop at a rate of about 100 per minute and grind about 3 to 5 tons of rock per stamp per 24 hr. Water consumption averages ar~uncl 0.5 tons per ton of ore. The stamp mill was originally devised as a combination grinding and amalgamating device before the days of cyanidation. Its lise continued with the introduction of the cyanide process, where it was well suited to the comparatively coarse crushing used, the distribution of the ground pulp over amalgamation plates, and the steps of separate cyanidation of sand and slimes that followed. As the "all-sliming" method became more generally adopted, however, with the need for fine grinding in ball mills and preferably in cyanide solution, the stamp mill tended either to be used as a secondary crusher or to be replaced altogether by dry-crushing equipment. Small stamp-mill installations are still to be found, and a number of large stamp mills are still in operation on the Rand, but no large new mill has installed stamps in the last 20 years. Chilean and Huntington Mills. These two types of mill are practically obsolete. In these mills rollers driven from a central gear-driven spindle revolve around a pan. In the former the rolls crush against a ring in the bottom of the pan, and in the latter centrifugal force holds the rollers against the ring at the side of the pan. Chilean mills were used at the Golden Cycle up to a few years ago for grinding roasted ore.

CHAPTER VI

Classification The metallllrgical advantage of fine grinding in the treatment of many ores by the all-sliming cyaniding process has of cow'se been realized ever since the introduction of the process. An economical means, however, of attaining a finely ground product has been a gradual development to which improvements in grinding mills and classifier« have contributed the greatest advance. The mill-claseifier combination Sh01dd be considered as a unit. The change from open circuit to closed circuit, the use of optimum circulatiru; loads, and the development of secondary and tertiary grinding circuits are notable advances in produciru; an improved product, generally reslllting in a higher extraction of the vallwble constituent» of the ore. This has been accompanied by a reduction in unit costs throllgh a lowered ctmsum/ptioti of power and steel and by increase in capacity of the grinding mill. DEFINITIONS

Classification as applied to cyanide plants is usually a combination mechanical-hydraulic operation which separates the solid constituents of a flowing pulp into two portions according to their respective settling rates. Usually it implies the removal of a finished product, termed "overflow," from a product requiring further grinding, termed' 'sand." Pulp means a uniform suspension of finely divided solids in liquids as applied to mechanical classification, which is the usual method of making a sand-slime separation in cyanide plants. Overflow is the comparatively finer, more slowly setting portion of the original pulp which is carried over the tailboard or lip of the classifier by the flow of water; sand is -the comparatively coarser, more rapidly settling portion of the original pulp which is discharged from the classifier by the mechanical action of the rakes. Selective classification is classification that has for its object the concentration of the heavy constituent of the original pulp, generally the sulphide, in the sand product, so that it may be ground finer than the lighter portion, generally the gangue. A grinding mill may operate open circuit or closed circuit. Open-ciccuit grinding is a method of comminution that produces the desired reduction in particle size by a single passage of the material through a mill. Closedcircuit grinding is a method of comminution in which a partly finished mill discharge is separated by the classifier into a finished overflow product 74

CLASSIFICATION

75

and an unfinished sand product which is returned to the mill for further grinding. USES OF CLASSIFICATION

1. To make a size separation, for example, a sand-slime 'separation- so that each product may be given a different treatment. 2. For closed-circuit grinding, so that the maximum size of particle escaping from the circuit may be limited and so that the useful work done by the mill may be increased by enabling the mill to operate largely in reducing oversize rather than in overgrinding majerial already fine enough.

FIG. 19. DOlT FX (heavy-duty) classifier with eccentric and rocker-arm mechanism.

3. For differential grinding of the heavy mineral constituent and the gangue, by selective classification which. concentrates the mineral in the sand product and causes it to be ground finer than the gangue before it can escape in the overflow product. The desired function, together with the character of the ore and such factors as the tonnage to be handled, mesh of separation, etc., determines the selection of the proper size and type of classifier. The single-stage classifier such as the DOlT (see Fig. 19 and 20), Akins, and Wemco, and bowl-type classifiers such as the DOlT (see Fig. 21) are in use for making separations from as coarse as 20 to as fine as 350 mesh. The first of these are better suited for making coarse separations and for handling large circulating loads where relatively small' overflow capacity is required. The bowl classifier, limited to 48 mesh and finer, with the double washing action gives a closer separation at the desired mesh and a cleaner sand product;

76


permits, through the selection of the proper bowl size, any required relation between overflow and sand-raking capacity; and is more suitable as a mineral concentrator for selective classification.

FIG. 20. DOlT HX (heavy-duty) classifier with slide arms and hydraulic lift.

FIG. 21.

DOlT

turret-bowl classifier installation in a large Canadian mill.

Both machines are capable of adjustment to cover considerable range of conditions. Adjustments in slope of classifier tank, rake speed, and dil ution of overflow pulp and, in the case of the bowl classifier, additional adjustments of backwash water and bowl speed are available to obtain efficient

CLASSIFICATION

77

results at any mesh of separation. In designing the machine, due consideration is given the results desired under specified condi tions; final adjustments are made when the particular flow sheet is placed in operation and are varied thereafter as the conditions vary. The evolution of fine-grinding flow sheets can be described as follows: Open-circuit Grinding. The first application of the mechanical classifier was to open-circuit work with grinding mills in the cyanidation of gold and silver ores, for the purpose of producing two products-a slimefree sand for treatment by percolation and a sand-free slime for treatment by agitation and filtration or clecantation. Closed-circuit Secondary Mills. Advances in the art of treating cyanide slimes and the reduced cost of fine grinding made possible by the early tube mills led to the use of the mechanical classifier as a means of controlling the fineness of the pulp leaving the grinding plant as feed to the all-slime cyanide treatment. Operating in closed circuit with the tube mills, the classifier not only controlled fineness more accurately but greatly reduced grinding costs. Closed-circuit Primary Mills. The adoption of two-stage fine grinding, due to a recognition that too great a size reduction in one mill with one ball charge was uneconomical, led to the use of the primary mill and subsequently a mechanical classifier in closed circuit with it. Reductions in grinding costs were again obtained, and in addition a convenient means was provided for apportioning the work between the two stages. Intermediate' Classification between Stages of Fine Grinding. Another step was the interposition of a bowl classifier between the primary and secondary grinding circuits. The classifier operated in open circuit receiving as its feed the overflow from the primary circuit, overflowing material of finished size, and discharging a clean sand product direct to the closed-circuited secondary mill. This step, however, is not generally desirable largely because of the difficulties met in controlling the succeeding classification stage in the absence of primary slime. Finishing Bowl Classifiers. Finishing the entire mill feed in a bowl classifier, with provision for regrinding the sands in a separate or tertiary circuit, first proved atthctive in the cyanidation of gold ores in which the values were chiefly associated with the heavy pyrite constituent. Here the bowl reclassified selectively the relatively coarse pulp from the secondary circuit, overflowing the bulk of the quartz and only the finest sulphides for treatment, concentrating the sand product to five or six times the assay of the original ore, and regrinding this high-grade material as fine as its assay warranted. The net result has been a higher extraction, a coarser over-all grind, and a reduced grinding cost.

78

CYANIDA'l'ION AND CONCENTRATION OF ORES

The same principle has been applied successfully more recently in the dressing of copper sulphide ores for flotation. The sulphide mineral, being heavier than the gangue, concentrates readily in the bowl sand and is subjected to regrinding for more complete liberation of associated minerals. In this way the mill is not burdened with gangue material 8.

TABLE

CLOSED-CIRCUIT

No. I Headst

classifier sands

Mesh

"'t., %

No.1 B.M. discharge

GRINDING

No.1 clussifier overflow

ANALYSES*

No.2 B.M. discharge

Cum., Wt., Cum., Wt., Cum., Wt., Cum., Wt., Cum.,

%

-0.2525 in. 8.6 8.6 4 38.3 46.9 6 7.8 54.7 10 11.7 66.4 20 10.2 76.6 35 7.3 83.9 48 2.1 86.0 65 2.0 88.0 100 1.7 89.7 150 1.6 91.3 200 1.1 92.4 -200 7.6 100.0 Per cent solids ...... Specific gravity.....

-

%

%

%

%

%

%

%

%

No.2 classifier sands -,

ws., %

No.2 classifier overflow

Cum., Wt., Cum.,

%

%

%

-- -- -- -- -- -- -- -- -0.6 0.6 0.7 0.7 16.5 17.1 6.6 7.3 6.9 24.0 3.4 10.7 14.0 38.0 7.6 18.3 0.3 0.3 0.1 0.1 0.2 0.2 20.0 58.0 14.9 33.2 6.2 6.5 1.8 1.9 3.5 3.7 17.1 75.1 18.3 51.5 17.3 23.8 11.6 13.5 16.7 20.4 6.4 81.5 8.5 60.0 9.7 33.5 11.8 25.3 13.9 34.3 0.1 0.1 4.1 85.6 6.7 66.7 4.7 38.2 20.2 45.5 21.2 55.5 1.6 1.7 2.9 88.5 2.6 69.3 7.8 46.0 18.5 64.0 17.8 73.3 7.0 8.7 2.6 91.1 5.6 74.9 8.9 54.9 17.3 81.3 15.4 88.7 14.7 23.4 1.5 92.6 3.9 78.8 6.6 61.5 7.0 88.9 6.0 94.7 12.8 36.2 7.4 100.0 21.2 100.0 38.5 100.0 11.2 100.0 5.3 100.0 63.7 100.0 82.1 2.81

78.2 2.87

52.2 2.78

70.4 3.04

77.4 3.01

15.8 2.81

Mesh . 20 35 48 65 100 150 200 Average 124 No.1 classifier circulating load. " . 108 117 123 151 121 124 125 No.2 classifier circulating load .... 360 346 372 365 402 427 436 387 • Mill of the Cariboo Gold Quarta Mining Co. (August, 1941). Samples taken on Aug. 27, 1941, over 24-hr. period. I t Heads: Per cent moisture. . . . . . . . 3.00 Specific grllvity.......... 2.86 Ore milled 373.0 dry tons ~

upon which no further work is required. Moreover, the finishing-bowl classifier protects the flotation operation against tramp oversize from preceding classifiers and assures a uniform product from the entire finegrinding section, convenient for sampling and distribution to the succeeding units. MEANS OF CLASSIFICATION

The type of machine best suited to a specific classification problem will . depend upon (1) whether a finished product is required, (2) the size range

CLASSIFICATION

79

of material to be handled, and (3) the volume and dilution of the pulp entering the machine. In general, open-tank rake-type, screw-type, or bowl classifiers are used for closed-circuit grinding work: where a finished product is desired. In some cases, however, especially for very fine separations requiring high overflow dilution, hydroseparators are used to good advantage, and because of the difficulty of raking extremely fine sand up an inclined deck, bowl-type classifiers have been converted to shallow hydroseparators with spigot discharges. The open-tank rake-type or screw-type machines are also more generally suited to coarser sizes of feed, while the bowl classifier, hydras, and centrifuges are applied more generally to handling finer (minus 20 mesh) feeds and to making finer separations. Where exceptionally large volumes must be handled, however, the hydroseparator has been installed to handle feeds as coarse as 4 mesh, using a spigot discharge. Hydraulic classifiers operate in the size range of about 8 to 200 mesh but find their principal use in preparing a number of closely sized products and in the concentration of heavy minerals. CONES

Classifying and dewatering cones once had a wide vogue and still are used in some pulp circuits. They were more or less troublesome because solids tended to accumulate on the sloping Sides, sloughed off periodically, and plugged the discharge in the apex; the sand discharged was too dilute and contained too much slime; and sand was frequently carried over with the slime. Furthermore,large cones required considerable space and headroom. It is of.interest that the Homestake, long a user of cones for dewatering and classifying, selected mechanical bowl classifiers for their new mill addition. In the Homestake's enlarged and rearranged plant are 10 cones of local make, 7 ft. in diameter and with 65-deg. sides. They dewater the pulp from the stamps and feed 10 rod mills which are in closed circuit with an equal number of Clark-Todd amalgamators and Dorr classifiers. Cones are used in the Dome plant, Ontario. There are two distribuiting cones and 16 classifying ones, 47~ ft. deep, with 60-deg. sides and %-in. spigot, which tend t» build up sulphides in the circuit for selective grinding. Cones of the Caldecott type, fitted with automatic discharge, are used in four plants at Kolar, India, for dewatering and making tube-mill feed. In this cone the discharge continues to operate until there is a change in the pulp density. If the pulp in the discharge becomes too thin, the plug at the cone bottom closes, allowing the sand to accumulate. The density then increases to the desired point after which automatic control becomes effective.

80


Cone classifiers in Rand stamp mills consist of a nest of cones 4 ft. in diameter and 5 ft. deep. The overflow passes to secondary cones to be separated into sand and slime. The underflow is divided into as many portions as there are tube mills, each portion gravitating to a dewatering cone 572 by 772 ft., the underflow of which is the tube-mill feed. A little water is added to bring the moisture content to 30 per cent. The cone, however, is inefficient, according to T. K. Prentice in Bul., I.M. and 1J1., April, 1935, in dewatering a finely ground pulp for tube milling. It has been found that the Dorr classifier is preferable for this purpose, and this machine is included in the flow sheets of all of the latest plants.

FIG. 22. Akins type S spiral classifier. MECHANICAL CLASSIFIERS

The impossibility of preparing clean, leachable sand and sand-free slime from an irregular feed of varying ore by double-cone classification led to the invention of the original Dorr classifier at the Lundberg, Dorr, and Wilson mill at Terry, S.D., in 1904. Mechanical dewaterers had been used before this time, including Johnson's shovel wheel, Scobey's belt dewaterer, and other devices, such as George Moore's cylinder with spiral conveyor. The Akins ribbon screw followed about 1910, and Philip Argall's double solid spiral, the Avoca, a few years later. The Scobey belt was developed at the Esperanza and used especially in Mexico, and later a cylinder with spirals introduced by Hardinge, while on the Rand a spiral draining sand from a cone has met some favor. Screw Classifiers. Screw classifiers such as the Akins and Wemco consist of a semicylindrical trough, usually set at a slope of from 272' to 372' in. per ft., in which rotates a helix or spiral at from 2 to 8 r.p.m.depending upon size and conditions to be met. The feed is introduced through the side of the tank just below the pulp

81

CLASSIFICATION

level. The heavier solids settle out and are carried upward out of the pulp by the spirals and discharged into a sand launder, while the fines flow in the opposite direction to the overflow weir at the lower end of the .

~~

The earlier types of machine used interrupted spirals and overflow Iips below the shaft carrying the spirals at the lower end. The design has now been modified. In the "high-weir" type recommended for coarser separations the spirals extend above the pulp level, and in the "submerged" type recommended for fine separations the spirals are below the pulp level in

FIG.

23. Wemco S-H triple-spiral classifier.

the lower end of the tank. Either type may be supplied with single or double spirals, and both are equipped with lifting devices. Figure 22 shows an Akins submerged-type machine. In the triple-pitch Wemco design (Fig. 23) three spirals per shaft provide additional sand-conveying capacity per revolution. Reciprocating Rake Classifiers. The DOlT Company is the principle advocate of this machine. Single-stage Type. The DOlT classifier (Figs. 19 and 20) consists of a rectangular, inclined settling box of wood, steel, or concrete, with the upper or discharge end open, in which are placed mechanically operated rakes or scrapers which carry the quick-settling, coarse particles to the point of discharge at the open end. Each rake, of which there may be as

82

CYANIDATION AND CONCEN'l'RATION OF ORES

many as six, is carried by two hangers, one at the discharge end and one near the overflow end. The rakes are raised, lowered, and moved parallel to the sloping tank bottom by a suitable head motion. A lifting device is provided for raising the rakes several inches to clear the settling solids in case of a shutdown. The feed enters continuously, through a distributing launder near the overflow end. The more quickly settling particles fall to the bottom and are advanced up the inclined deck by the rakes and discharged. The agitation of the reciprocating rakes keeps fine particles in suspension until they overflow the weir at the lower end. The point at which a separation can be made is determined by the rake speed, the pool area, and the overflow dilution. The greater the rake

FIG. 24. General arrangement of Hardinge countercurrent classifier.

speed the greater the overflow density, and the smaller the pool area the coarser the separation. Bowl Type. The Dorr bowl classifier (Fig. 21) is a combination of a shallow, circular bowl with a revolving raking mechanism superimposed on the lower or overflow end of a single-stage Dorr classifier. Feed enters through a loading well at the center of the bowl, and fine solids overflow across a peripheral weir. Coarse solids settle on the bowl bottom, are raked to a central opening, and gravitate through it into the reciprocatingrake compartment. It is used where a clean rake product is desired, where a separation at a fine mesh is to be made, or where the overflow capacity must be large in comparison with raking capacity. Hardinge Classifier. The Harclinge Company was responsible for introducing the spiral ribbon type of mechanical classifier as illustrated in Fig. 24. This classifier is a slowly rotating chum, on the inner surface of which is a screw flight attached to the drum, revolving with it. The material to be classified is fed in at one end above the pulp level, and as the

CLASSIFICATION

83

classifier rotates, the coarse particles that settle out are moved forward by the screw flight. The fines overflow through an opening at the feed end of the classifier. The sand or oversize is dewatered and elevated by buckets to the discharge hopper. HYDRAULIC OLASSIFIERS

Hydraulic classifiers range from simple V-shaped launders with a multiplicity of shallow settling pockets for the discharge of as many roughly sized products to the more elaborate deep-pocket machines of the hinderedsettling type, having specially designed constriction plates and automatic discharge of spigot products. Dorrco Sizer. In the eight-pocket machine, the pockets are separated by 3-ft.-high partitions and increase in cross-sectional area from 1 sq. ft. in the feed end to 3.12 sq. ft. in pocket 8. The size of hole and hole spacing in the constriction plates depend upon the size distribution of the feed and are designed individulally for each installation, so that the hydraulic water required for each compartment will flow through the holes at a predetermined head loss. This hydraulic water is supplied from take-offs arranged along a common manifold, and the flow is controlled by individual valves. Windows are provided in each compartment for inspection of the teeter bed. The discharge mechanism is fully automatic, the spigot valves being operated by a modulating motor controlled by a pressure unit that, in turn, is actuated by the pulp density within the compartment. This eight-pocket unit operating on minus 14-mesh iron ore of average size distribution has a feed capacity, when overflowing minus 100-mesh material, of about 25 tons of solids per hour. Hindered settling machines require no power except for the water used and are automatic in discharge, although they need a certain amount of attention. On many materials they give excellent results either in closeness of sizing or in concentration of the denser particles. They are used, therefore, chiefly for two purposes-either in separating 6- to 200- mesh material into a series of closely graded portions for subsequent treatment or in concentrating the heavier mineral of a nonhomogeneous feed. In general, however, hydraulic classifiers use more water than mechanical classifiers, the slime overflow is more dilute, the coarse fractions contain a much higher moisture content, and the loss of head on the sands is considerable, which is objectionable in closed-circuit grinding. Also, these classifiers frequently require a deslimed feed. Thus in practice, for preparing cyanide or flotation pulps, mechanical, nonhydraulic classifiers are used almost to the exclusion of all other types because of their capacity, flexibility, continuous operation, and uniformity of results. In addition, they require little attention and give a coarse

84


portion or sand of low moisture content. Furthermore, the mechanicaltype classifiers give a coarse portion that is sufficiently free of fines for practical closed-circuit grinding. This is the result of the agitation and draining that take place during the removal of the coarse portion. }IYDROSEPARATORS

A hydroseparator is, in working principle, an undersize thickener, i.e., a machine of such a specified diameter that the upward displacement rate is greater than the settling velocity of the largest particle it is desired to separate. In consequence, particles of this critical size and finer are carried into the overflow, and the settled material, which is carried to a central discharge opening by a slowly moving raking mechanism, is relatively dislimed. As in the case of mechanical, in contrast to hydraulic, classifiers generally, a considerable amount of the finer fraction is entrained with the underflow sands. Used both as "deslimers" for making final separations after grinding and also for closed-circuit grinding work, they have the advantage of providing the relatively large areas required for fine separations at high dilutions. Where it is necessary to handle heavy sands in the machine, steep sides and extra-heavy raking mechanism are provided. The capacity of hydroseparators can be determined from the following formula:

A

1.333 =

(F + sp._1_) gr. R

where A = square feet per ton of overflow per 24 hr. F = overflow dilution R = settling rate in feet per hour at over-flow dilution F sp. gr. = specific gravity of the solids Having determined the ratio of feed to overflow, the unit area per ton of feed is readily calculated. DESLIMING

Desliming for Separate Treatment of Primary Slime. A number of mill flow sheets include the step of removing the softer fraction of the ore (clay and related material) from the harder fraction preceding or as a part of the crushing and grinding sections. At Marlu Gold Mining Areas in West Africa (Fig. 89) the ore after passing through a jaw crusher is washed at 17~-in. size in a series of twelve 4 by 16-ft. trommel washers, the undersize being sent to bowl classifiers for

CLASSIFICATION

85

desliming and the washed rock being crushed in Newhouse crushers and passed to the ball-mill circuit (which also receives the sands' from the desliming classifiers). At the Dome mines in Canada (Fig. 64) the jaw-crusher product is first washed on vibrating screens, the oversize passing to a, Symons cone crusher, while the undersize is classified for removal of fine sand and slimes and the rake sands transferred to the mill bins. In both of the above installations the objective is, of course, to eliminate primary slime which in the case of wet ores causes trouble in the crushers, but the' combined pulp is cyanided in a single circuit. In the case, however, of the treatment scheme devised to handle a difficult ore in the plant of the Kelowna Exploration Co. at Hedley, British Columbia (Fig. 71), the primary slime and softer fractions of the ore are treated in a separate circuit from that used to handle the harder ore constituents. This scheme has resulted in the successful treatment of an ore that otherwise failed to respond to ordinary thickening and filtering practice. CENTRIFUGAL CLASSIFICATION

Bird Centrifuge. This machine is essentially a closed spiral classifier in which the tank is revolved at relatively high speed on a horizontal axis. The tank itself consists of a truncated conical shell, within which is a smaller concentric cone carrying a spiral ribbon (rakes) that revolves independently of the shell but in the same direction at a somewhat slower speed. Classification takes place in the annular space between the cone and the shell, the speed differential between the two elements having the effect of raking the solids settled against the shell up-slope to the small end, where they are discharged. The slime discharges through ports at the larger end. The machines are built in various sizes ranging from a raking capacity of 2 to 50 tons per hour, the latter requiring a 100-hp. drive motor. At the Hedley Mascot mill in British Columbia, a change in the ore mined led to a serious slime problem which could not be handled in the conventional flotation and cyanide equipment installed at the mine. Following test work, a 36- by 50-in. Bird centrifuge was installed in January, 1941, to deslime the ore prior to flotation and has been in continous operation since. According to C. W. S. Tremaine in "Applications of the Bird Centrifuge at Hedley Mascot Mill," C.I.M. and l11I., Vol. 50, pp. 533-536, 1947, the centrifuge is operating on a mixture of slime from primary and secondary classifiers and concentrate taken from the scavenger flotation cells. This mixture of slime and low-grade concentrate forms the feed to the Bird centrifuge; it amounts to about 2.5 tons per hour at 15 per cent solids, 98 per cent minus

86


325 mesh, and assays about 0.11 oz. per ton. The centrifuge discharges an effluent which carries 3 to 6 per cent solids and assays 0.06 oz. gold per ton. The cake is discharged at about 82 per cent solids, 92 per cent minus 325 mesh, and is sent to the cyanide plan t, where it is treated in conventional manner. . .. The centrifuge machine is driven by a 30-hp. motor and operates at 1000 r.p.m.

D.S.M. Cyclone. The hydraulic cyclone, which was developed by the Dutch States mines in connection with their coal-cleaning processes, is the most recent classifying-thickening device to be used commercially for the separation of fine particle sizes. The design and principle of operation of the hydrocyclone are similar to that of the familiar dry cyclone or dust' collector. It is a closed vessel consisting of a conical section surmounted by a cylindrical section of equal diameter. The feed is pumped into the cylindrical section at a pressure of 5 to 50 lb. per sq. in. through a tangential opening, the fines overflowing through a central orifice on top and the coarser fraction discharging through a second orifice at the apex of the cone. The high rotational velocities developed inside the vessel set up centrifugal forces equal to many times the force of gravity, and high rates of separation are obtained. The coarser material is thrown to the sides of the cone and forced by a pressure differential toward the discharge orifice, emerging as a thickened sludge, while the fines are displaced through the overflow opening. The capacity of thehydrocyclone and the separations made depend upon a number of design factors, including the shape and size of the vessel, the pressures used, and the size of overflow and discharge orifices. Though still inrthe experimental stage, performance data to date indicate that this device will find considerable use in the field of classification and desliming. It possesses the advantage of high capacity for size and relative simplicity of construction, which involves no moving parts. The only power required is that used for pumping the feed into the vessel.

CHAPTER VII

Sand 'Treatment Ores may be separated into sand and slime, and the sand leached by upward or downward as well as upward and downward percolation of cyanide solutions. Leaching practice has decreased in. recent years with the development of closed-circuit grinding and now is seldom incorporated in new plants. Ores are known which need only to be crushed coarsely to 7~ in., and the whole leached. Deposits of ore amenable to coarse crushing, however, are rare. The gold must be on cleavage planes or in the cavities left by previous natural leaching of the ore in place.

Sand leaching is carried out in vats ranging in capacity from 30 to 1200 tons. Sand for leaching is separated from slime in cones, V boxes, classifiers, and in collecting vats filled by distributors-the overflow in each case being slime or finer portion of the ore. As a rule, leaching is a simple process, involving a vat of well-mixed neutralized sand, ample contact with strong and weak cyanide solutions, water washes, aeration of the sand and solution, and rapid filling and discharging. It is a cheap and effective process for clean ores when fine grinding is not necessary for good extraction. Classification for Leaching. The importance of classification prior to leaching cannot be overemphasized. Sands that are essentially free from colloidal material behave quite differently from the same type of sand with a small percentage of colloid. The two following examples from plant practice are illustrative. The Golden Cycle mill at Colorado Springs, Colo., grinds roasted siliceous ore in rod mills through 20 mesh before sand-slime separation at about 200 mesh. Prior to the development of the Dorr bowl classifier at this plant, this separation was made in Dorr classifiers which produced a sand containing about 15 Pel: cent minus 200 mesh, with an appreciable amount of contained colloid. The leaching rate in the same tanks varied from 0.7 to 1.4 in. per hr. Following the adoption of the bowl classifiers, which produced practically colloid-free sand, the leaching rate increased to about 10 in. per hr., and leaching results were much more uniform. The table on the following page shows a typical screen analysis of the bowl-classifier sand when the overflow contained 2.6 per cent plus 200 mesh. 87

88


+

30 + 60 +100 +150 +200 -200

mesh mesh mesh mesh mesh mesh

6.7 - 53.2 - 73.2 - 84.2 - 96.0 - 100.0

cumulative cumulative cumulative cumulative cumulative cumulative

The Homestoke Mining Company grind their ores in closed circuit with rod and ball mills through about 80 mesh and, following amalgamation, make a sand-slime separation at about 2 per cent plus 200 mesh. In their new plant, DOlT bowl classifiers make this separation prior to leaching and slime treatment. The Homestake ores contain an appreciable amount of pyrite which is gold bearing, and the bowl classifiers are adjusted to produce a maximum of minus 200-mesh granular material in the sands, including very fine pyrite, but to remove all colloidal material. Even TABLE

9.

SIZING ANALYSIS OF SAND AT THE I-IOlvIESTAKE

(East Sand Plant) Mesh

+ 50 + 80 +100 +150 +200 -200

Per cent

Per cent cumulative

0.1 3.0 8.0 17.5 25.0 46.4

0.1 3.1 11.1 28.6 53.6 100.0

with this unusually fine sand, excellent and uniform leaching results are obtained at rates up to 2~/z in. per hr. in 11-ft.-depth tanks. Table 9 is a typical screen analysis of this sand. Homestake Leaching. Although it is finely ground in water, Homestake sand can be cyanided by gravity leaching to yield a high extraction of the gold. Sieve analyses and assays reveal that the finer the sand the higher the extraction. This varies from better than 90 per cent in the minus 200-mesh material to less than 50 per cent in the case of particles coarser than 50 mesh. Leaching is done in two sand plants, one with 8 vats 44 ft. in diameter by 12 ft. deep, the other with 21 vats 44 ft. in diameter by 11 ft. deep. They are filled through Butters-Mein distributors. Lime is added to the pulp as it leaves the final set of classifying cones. The filter bottom consists of the usual wooden frame covered with coco matting and 10-oz. canvas duck. Repeated aeration is essential to maintain enough oxygen in the solutions to effect extraction, according to A. J. Clark in E. and 111.J., Oct. 12, 1931. This is done by introducing low-pressure air under the filter bottom of the leaching vat.

89

SAND TREATMENT TABLE

10.

SAND-TREA'l'MENT CYCLE AT

Period, hr.

Operation

Filling

.

10

First draining First aerating

. .

20

16 to 28

First solution leaching ....

16 to 24

Second draining,

.

14

Second aerating

.

6

Second solution leaching ..

14 to 16

Third draining

.

14

Third aerating .. ,

.

6

Third solution leaching ...

14 to 16

Fourth draining

.

14

Fourth aerating Fourth leaching Washing

,. . .

Sluicing........ .. .. .... .. . Total

.

THE I-IOMESTAKE

Remarks

Pulp containing 43% solids joins discharge of lime mill and enters' waterfilled vat through Buttcrs-Mein distributor. Overflow, after clarification, is reused in classifiers. Effluent wasted. Gage pressure, 7H lb.: bottom solution valves clpsed. Time of aeration depends on alkalinity. Rate, 2 in. per hr.; solution strength, 0.095% N aCN; displaced moisture wasted. Effluent wasted. From appearance of gold, solution sent to precipitation. Gage pressure, 8 lb.; bottom solution valves closed. Time depends on flow rate, approximately 2 in. per hr.; solution strength, 0.09% N aCN; effluent to low-solution sump, followed by precipitation; barren solution reused as wash. Effluent to weak-solution sump, followed by precipitation; barren solution to storage for reuse. Gage pressure, 8 lb.: bottom solution valves closed. Solution strength, 0.055 % NaCN; effluent to weak-solution sump, followed by precipitation, and discharge of barren solution to storage for reuse. As for third draining. Barren solution 0.035% NaCN. Effluent to precipitation for first 16 hr., then to strong solution make-up.

6

24 to 28

18

5 197 to 225

hr. 8 to 9 days

Rate, 2>~ in. per hr. Effluent to strongsolution sump for reuse. Wash water from cone overflow. About 40 to 48 in. of wash water used.

90


The leaching practice is designed to effect a progressive enrichment of the solution before precipitation, the effluent from the latter part of a treatment being strengthened but not precipitated before being returned to the top of another charge. Solution and zinc are conserved by this system. The total solution used amounts to only 0.8 ton per ton of sand. The water draining from the newly filled charge contains sulphates and thiosulphates. After the second aeration, when cyanide solution has been added to the top, these compounds appear in larger quantity. Later, they are followed by thiocyanates, which in turn are followed by the first traces of cyanides. Gold appears in the effluent soon after the first trace of cyanide is noted. Thiocyanates, the main source of cyanide loss, are TABLE

11.

SAND-SLIME SEPARATION AT THE GOLDEN CYCLE

Percentage

Tyler standard size Feed

On l4-mesh screen .......................... On l6-mesh screen .......................... On 20-mesh screen .......................... On 30-mesh screen .......................... On 40-mesh screen .......................... On 60~mesh screen .......................... On 100-mesh screen ......................... On l50-mesh screen ......................... On 200-mesh screen ......................... Through 200-mesh screen ...................

0.2 0.9 5.7 19.8 30.2 53.2 56.7 63.3 69.3 30.7

Moisture .................................

86.7

cum. cum. cum. cum. cum. cum. cum. cum. cum. indo

Sand

Slime

0.2 cum. 0.8 cum. 7.4 cum. 33.7 cum. 52.8 cum. 85.3 cum. 92.7 cum. 98:0 cum. 99.4 cum. 0.6 indo

0.8 3.5 15.9 29.1 70.9

23.3

93.8

cum. cum. cum. cum. indo

stable compounds, but although they build up in solutions, they do not affect extraction. Low alkalinity is favored, a pH of 9.6 to 10.2 giving the best extraction. A cyanide of 49 per cent N aCN equivalent is used. Strong solutions carry 0.095 per cent and the effluent 0.035 per cent NaCN. Cyanide consumption is ;!.i lb. per ton treated. Golden Cycle Leaching. At the Golden Cycle, Dorr and Akins classifiers make a sand-slime separation as shown in Table 11. The sands are conveyed to 10 leaching vats 50 feet by 15 ft. deep with a capacity of 1200 tons each. The initial leaching period is 48 hr. with solution containing 0.5 lb. cyanide per ton. This solution goes to the precipitation presses. Sand charges are drained and aerated at least four times for 8-bJ... periods during the 6- to 8-day treatment cycle. Barrensolution washes and a water wash follow. Two men using 2':>'2"-in. hose and water at 120-lb. pressure discharge the vats in 37'2" hr.

SAND TREATMENT

91

SAND LEACHING ON THE RAND

While the trend on the Rand is toward all-slime plants, a considerable tonnage is still handled by leaching of the sand fraction of the ore. Randfontein Estates Gold Mining Co. This mill, i~hich is described in detail in Chap. XV, Sec. 3, mills 13,000 tons pel' day, of which 20 pel' cent is handled by sand leaching. -Aera/ed .so/or/on ==-~---=:::=.--~-~~~

!~f/ifi~~

Pressvre goge /0 /b./s'l-li?

/

5 "Leaching pipe

.st/o-aerar/or:

Leachingpipe

FIG. 25. Aeratiun of solution with sub aeration and vacuum drainage of sand charge.

The sand from bowl classifiers at 51.8 pel' cent plus 100 mesh, 8.5 pel' cent minus 200 mesh, is pumped to Butters distributors serving thirteen 50-H.-diameter by 12-ft.-deep sand-collecting tanks, which, in turn, discharge the collected sand to twenty-six 50-H.-diameter by lO-ft.-deep leaching vats. Treatment cycle is 22 hr. collecting and transfer, 10 hr. leaching and washing, and 4 to 5 hr. emptying. A charge of about 1000 tons of sand is leached with 250 tons of strong solution made up to 0.055 per cent KCN, followed by 750 tons of plant circulating solution and then 1000 tons of barren solution. Seventy-four per cent of the effluent at an average value of 1.5 dwt. per ton goes to precipitation. The balance is by-passed to plant circulating solution.

92

CYANIDA'rION AND CONCENTRATION OF ORES

LEACHING OF CLAY ORE

Bidi, Sarawak, Borneo. To solve the problem of treating a clay ore in which the gold was finely disseminated, the author was able to suggest a unique method 'which he had seen described twenty-five years previously in a paper, "Occurrence and Treatment of Gold Ore at Bidi, Sarawak, Borneo," by T. C. Scrutton, which appeared in Trans. 15, I.11iJ.Jl!I., 19051906, in which clay ore was treated by leaching in relatively large lumps. Its successful application in this instance is our justification for including a detailed description of the method in the present volume. The auriferous deposits at Bidi consist of a series of unconnected hills lying upon the weathered surface of limestone. These hills consist of clays and earth containing boulders of stone of varying structure-pure silica, silicified sandstones and shales. The clays and earths carry from 2 to 15 dwt. gold per ton, averaging 5 dwt.; the stone, from 3 to 30 dwt. per ton, averaging 8 dwt. The proportion of stone to clay occurring in the deposits is roughly 1 to 5. The ore is trammed to the six ore bins. The clay ore bins are four in number, and their arrangement constitutes an important feature in the direct treatment of clays. To quote Scrutton: To obtain rapid and easy leaching of the clayey material, it is necessary 1. To keep the fine earthy and sandy material separate as far as possible from the more plastic clayey material and to treat the two separately in different vats. 2. To ensure that all earthy and clayey material, when charged into the vats, is in the form of balls firm enough to maintain their form when charged into the vat and of a size varying from 3 in. in the case of the clayey material to 7i in. in the case of the finer. The first of these conditions is attained by providing four separate clay bins and carefully selecting the material from each according to its tendency to break up and form fines or to agglomerate and form balls. . The second is provided for by running the clay from the tippler down to a fanshaped chute, about 40 ft. long and inclined at 60 deg. into the ore bin; thus the masses of clay are broken up and formed into balls by rolling down the chute. The wetter and more plastic the clay treated the longer and steeper must this chute be to ensure the clay's being in a leachable condition on arriving at the ore bin. The clay ore bins are rectangular and discharge through bottom sliding doors into wagons running below. Vats. The leaching vats are 30 in number, 18 of 100 and 12 of 50 tons' capacity; the former are 6 ft., and the latter 3 ft. deep, all of 27 ft. 6 in. diameter. They are constructed of %-in. mild steel plates, riveted with H-in. rivets, 1%-in. pitch. Charging. Side-tip wagons are employed and are filled direct from the stone and clay ore bins, whence they are run direct over the vats and tipped; two pairs of rails run over each vat, arranged at such a distance apart as to require a minimum amount of shoveling to level off the material in the vat. The following are the principles regulating the method of charging:

93

SAND TREATMENT

1. Clay must be charged in layers not more than 3 ft. deep; if this depth be exceeded, difficulties are experienced in leaching and washing, resulting ·in impaired extraction. 2. Coarse material must be kept separate from fine to ensure good leaching. 3. In the event of charging two different classes of stone and clay 'into the same vat, the operation must be conducted so as to leave the material as far as possible in horizontal, not vertical, layers. . In the case of the 50-ton vats, which are only 3 ft. deep, they are filled with clay from one ore bin, no further precautions being necessary. With the 100-ton vats, in order to treat as large a quantity of clay as possible, sufficient clay is charged into the bottom of the vat to form a layer 2 ft. deep; this is then leveled off, and sufficient stone charged to form a layer 1H ft. deep; this having been leveled off, the vat is filled to the top with clay.t'giving an upper layer of 2 ft.; thus the layer of stone, by forming a porous bed in the middle of the clay, prevents the formation of channels throughout the whole mass and, by separating the clay into two thin layers, renders leaching comparatively easy; by charging in this manner equally good extractions are obtained from the clay treated in the large as in the small vats. To ensure obtaining the correct tonnage, it is necessary to fill the vats until the clay stands about 2 in. above the top of the vat, as the material when properly charged lies exceedingly loosely and, after solution has been on for a short time, sinks down to 9 in. below the top of the vat. Discharging. This is effected in the usual manner by shoveling the material through four bottom-discharge doors into wagons running on two lines of rails below the vats. Treatment. The usual treatment for 100-ton vats occupies 10 days and gives .the following results: Charge Discharge , Extraction KCy consumption

,

5.4 dwt. , 1.2 dwt. 78 per cent , 0.97 lb. per ton

The solutions used are: Strong solution, containing. . . . . . . . . . . . . . . . . . . . . . . . . . . . .. 0.07% KCy Sump solution, containing 0.05% KCy These solutions have to be kept at a definite standard of alkalinity, which is effected partly by putting a certain quantity of lime into each wagon at the bottom of the main incline and partly by the addition of soda to the sumps; the quantities of each used are regulated by the working of the zinc boxes. The average consumption of lime is 8.8 lb. per ton, and the maximum amount of soda usually added 0.14 lb. per ton of solution. So long as this standard of alkalinity is carefully maintained, the solutions give very little trouble. If, however, the alkalinity gets too high, the solutions become dirty, foul smelling, and full of arsenic and antimony; if too low, the consumption of cyanide is so great as to show no precipitate with AgN0 3 after once running through a vat, extraction, of course, suffering accordingly. The first filling of strong solution is put on by very slow upward leaching, so as to disturb as little as possible the fine material lying loosely in the interstices between the larger balls of clay. The vat is gradually filled by this means, the opera-

94


tion, if properly conducted, occupying about 3 hr.; as the charge in the vat becomes soaked in solution, it settles down, finally sinking to about 9 in. below the top. About 30 tons of solution is required for the first filling, but of this only about 20 tons can be drained off, the remainder being absorbed by the clay. The vat when full is shut up and allowed to soak for 4 hI'.; it is then opened and allowed to drain at such a pace as to ensure its being just dry in another 4 hr., when it is pumped up again, the solution being run on to the top of the charge, and the vat leached by gravity in the usual manner. This alternate filling by downward leaching, soaking, and draining is carried on until about 150 tons of strong solution has been put on the vat, i.e., for about four days. Sump solution is then substituted, the first filling being put on by upward leaching, and the subsequent filling by downward leaching, in the usual manner. The effect of using slow upward leaching in the middle of the treatment is to lighten the charge and form new channels for solution by altering the direction of pressure. The downward leaching is continued as before until about 180.tons of sump solution has been given. This usually occupies about 5 days; the actual time of treatment, however, is judged by the amount of gold extracted according to the solution assays. When the solutions coming away contain only H dwt. per ton of gold, a final water wash is given, and the vat discharged. In order to ensure a regular flow of solution through the zinc boxes, it is necessary to divide the vats into two lots, one lot being drained while the other is soaking. After repeated trials the foregoing method has been found to give the best results for clayey material charged direct into the vats. Given that the material has been properly charged into the vat, solution of the gold takes place almost immediately on contact with the fresh cyanide solution, the metal being in an extremely fine state of sub-division, and lumps of clayey material containing only about 16 per cent moisture being readily permeable by' solution. Practically, the total gold contents of the clays are dissolved by the cyanide; this was shown by taking a number of samples of the material discharged from the vats and applying repeated washings of water; by this means alone a final extraction of over 95 per cent was obtained. The problem, then, in order to ensure good extraction, is to wash out the auriferous cyanide which has been absorbed by the balls of clay. This cannot be effected by direct washing, in the ordinary sense of the term, the lumps of clay being only very slightly pervious and the interstices forming easy channels for solutions, but it must be brought about by diffusion between the solutions rich in gold remaining in the dissolved clay and the solutions containing practically no gold being pumped into the vats. In order to obtain the best results, the point to be aimed at is to give the charge as much fresh solution as possible, consistently with sufficient time of soaking to allow of a certain amount of diffusion between the fresh solution and the auriferous solution remaining in the clay; prolonged soaking has not been found to give good results; doubtless owing to the slower diffusion of liquids carrying nearly the same quantities of gold in solutions, as compared with those differing widely in gold values. There is, however, a decided limit to the amount of solution which can be used, as it has been found that, unless a sufficient time of soaking be given to allow the new solution to permeate the whole of the charge, the extraction suffers considerably, the new solutions coming away by certain easily formed channels and absorbing very little of the richer solutions contained in less readily permeable parts of the vat. Likewise slow draining off, i.e., at the rate of 5 tons solution per hour, is a necessity. H this rate be exceeded, a much larger proportion of moisture remains in the vat,

SAND TREATMENT

95

doubtless on account of the solution's descending too quickly into the vat to allow the small particles of solution lying in the interstices to agglomerate and descend with the mass of liquid; they are accordingly 'cut off by the air and left. Extractor House. The precipitation is effected in the usual manuerby zinc shavings, the zinc boxes being 16 in number and containing in all 936 cu. ft. zinc. At the rate of flow given this allows 1 cu. ft. zinc for 1 ton solution per 24 111:. • Considering the large amounts of antimony and arsenic contained in the charge, the precipitation gives very little trouble, though from time to time the solutions become abnormally foul and the precipitation is impaired. When working satisfactorily, 0.5 per cent cyanide solution entering at 2 dwt. per ton is reduced to 3 grains.

LEACHING AGGLOMERATED' SLIMES

More recent work along these lines is reported in T.P. 790, A.I.1I1.E.' by O. C. Shephard and C. F. Skinner presented at the New York meeting in 1937, under the title of "Stabilizing Agglomerated Slimes for Cyanide Leaching." The paper describes the development of a method of stabilizing agglomerated slimes by the formation of a cementing substance in the glomerules. The conclusions based upon agglomeration tests were as follows: 1. The porosity and permeability of finely ground ores can be greatly increased by moisture agglomeration. 2. Variations in the amount of solution used in agglomeration causes a noticeable difference in the amount of permeability. The maximum permeability is reached between 10 and 18 per cent moisture. 3. A point of saturation occurs when too much solution is added, causing the glomerules to break down to a runny mud. Beyond 22 per cent moisture, none of the materials tested had a measurable permeability by the method used. 4. The amount of solution necessary to give permeability by agglomeration increases with the fineness of the particles. 5. The permeability of loose beds of agglomerated material decreases with packing, but the permeability decrease becomes less as packing progresses.

CHAPTER VIII

Slime Treatment Slime treatment, as commonly ueed by metallurgical engineers, includes thickening, agitation, and filleTing and as applied to cyanidation also inciudee washing by continuous countercurrent decantation (C.C.D.) and/or filters. "Slime" is the general term used to describe the finer portion. of pulp in a combination sand and slime-treatment plant and is uS1wlly [iner than 100 mesh. The so-called "all-slime" type of plant is that in which all of the ore is qvound through a relatively fine mesh such as 100 mesh and wheTe no separate treatment of sand and slime is provided. Equipment used and methods employed aTe discussed, with examples [rom. practice. THICKENING

General. Thickening or dewatering may be defined as the removal of a portion of the liquid from a pulp or slime made up of a mixture of finely divided solids and liquids. The early methods of thickening employed plain, flat-bottomed tanks into which the pulp was fed until the tank was full. The solids were then allowed to settle as long as required, the top liquid was decanted, the settled solids were discharged, and the operation was repeated. Such settling was usually carried out in a number of tanks so that a regular cycle of filling, settling, and discharging could be maintained. Later it was found that feeding the tank behind a baffle allowed some decantation of clear liquid while still feeding, and this also was introduced. Attempts to make thickening continuous, by using hopper-bottom tanks or cones, were not entirely successful. In these tanks the feed ran in continuously, settled solids were drawn off through a spigot and solution overflowed continuously at the top of the tank. The chief drawbacks to this method were the multiplicity of units required and the. fact that settled solids hung up on the sloping sides which made it extremely difficult to obtain anything approaching a discharge of uniform density. Dorr Thickener. The invention of the Don thickener made possible the continuous dewatering of a dilute pulp whereby a regular discharge of a thick pulp of uniform density took place concurrently with an overflow of clarified solution. Scraper blades or rakes, driven by a suitable mechanism, rotating slowly over the bottom of the tank, 'which usually slopes gently toward the center, 96

SLIME TREATMEN'.r

97

move the material settled on the bottom to a central opening or discharge. The rakes revolve at a speed sufficient to move the material as fast as it settles without enough agitation to interfere with settlement. Dorr thickeners are used in the metallurgical field to thicken prior to agitation and filtration, in the countercurrent washing of cyanide slime, for thickening ahead of flotation, for thickening concentrates, and for dewatering tailing to recover the water for reuse in the mill. The standard construction of Dorr thickener mechanisms is of iron and steel. .The tanks are usually made of steel or wood for medium-sized machines, but in the larger sizes they are often constructed of concrete or earth or a combination of these materials. V~rious types of Dorr thickeners, which it is unnecessary to describe here, are available for specific uses, including constructions to resist corrosion. Power requirements are low, about 7~ hp. for thickeners less than 50 ft. in diameter. Attendance and repairs are also low. Figures 27 and 28 illustrate the development of the original central-shaft type of thickener from one using superstructure to the present beam type of support whereby headroom is saved and an improved mechanism running in oil is made conveniently accessible to the operator. The D01T tray thickener (Fig. 28) has been developed to meet the definite demand for large settling area in limited space. Each compartment increases the capacity approximately to the same degree as an additional thickener unit, of the same size, without increasing the floor space required. The tanks are divided into two or more settling compartments by means of steel trays or diaphragms suspended from the sides. The mechanism is made up of a central vertical shaft, driven by worm gear and with radial arms attached above each 'try. These arms carry plow blades set at an angle, and as the mechanism slowly revolves, they move the settled material to a discharge opening at or near center. Several types of tray thickeners have been developed, classified according to the arrangement of feed and discharge in the various compartments. Probably the most generally applicable is the balanced-type tray thickener. In this each compartment has a separate feed and overflow, but the settled solids from all compartments are brought together and are discharged through a central outlet from the bottom compartment. The Dorr traction thickener is the type most frequently used for heavyduty work with large tonnages. The machine is so called because the thickening mechanism is driven by means of a traction wheel which travels around the periphery of the tank on a rail. The mechanism, which is slowly carried around the tank as the traction wheel travels, consists of a truss to which are attached the raking blades which sweep the floor of the tame This truss is supported at the center of the tank by a column.

co 00

o

Overflow dischorqe one or two openinqs dependinq on the overflow

>-1 ".

Z H t:I ". >-3

sZ ;,.

Z

Drum

t:I

By pass outlet

II I f

II ,

liB',,"

II I '

II

f

II

f

II

f

II ' I I I , ~

Bypass inlet

1,1,1

~~==,;;-;i~:;"~fn~~~~-~~~~~-3~~~~~~;~~~;i~~ii~;;~~i~ ~~;J:J.hI\Lt.!:j@f

<::>

a

z

<::>

t'j

of pier

Z >-3 ~

". >-3 H

a

Z ,.. '3 dischorqe lines

SECTIONAL ELEVATION FIG. 26. Plan and sectional view of a Dorr traction thickener.

-1·1

Pump to be located to qive clearance from end of truss and bearinqs

aI=j

a

~ t'j

rJJ

99

SLIME 'l'REATMENT

Power is brought in through a conduit which comes up through tllis column and is carried out along the truss to the drive unit which operates the traction wheel (see Fig. 2G). The Dorr iorq thickener is a new.development in the unit thickener field. The raking mechanism is supported from a stationary central pier on which the driving motor and gear reduction are mounted. The feed may be introduced at the center through a launder suspended above the top of the liquid leavel or by means of a siphon feed through the hollow central pier. In addition to compact strength provided with the central pier construction, the machine is characterized by a new rake-lifting feature whereby when oJ

I

Ii 1

L___

_

__

__

__ ._~_. _ __

I i

_.J

FIG. 27. Original Dorr thickener as installed at the mill of the Mogul Mining Company, Pluma, S.D., 1906. (After patent drawing.)

overloaded the rakes raise backward and upward at an angle to clear the overload and at the same time maintain a full raking load until the overload is removed and the rakes assume their normal position. The Dorr washing thickener is of the multiple-tray type and is, in effect, a complete countercurrent decantation plant in a single compact unit. It is adapted to the needs of the small chemical and metallurgical plants where a relatively small tonnage of finely divided solids must be washed free from a solution in a minimum space. Four to five stages of washing are usually provided in separate tray compartments. Multiple-stage washing is effected by operating the tray compartments in series rather than in parallel. Feed enters, and strong solution leaves the uppermost compartment. Washed solids are discharged from the bottom compartment. Solids, after each successive settlement, gravitate from one

100


compartment to another through sludge seals in each tray. Wash water, introduced in the bottom compartment, rises successively through the compartments in a direction countercurrent to that of the sludge. One tall thickener thus does the work of several relatively shallow thickeners of the same diameter. Floor space is conserved, heat insulation is simple and effective, and power is reduced. Furthermore, all control of the washing operation is centered at a single point. Diaphragm Pump. The diaphragm pump is essentially one for low lifts and is particularly adapted to handle metallurgical pulps economically and with minimum attention. The Dorrco diaphragm pump was de-

FIG. 28. Type ATB,

DOlT

balanced-tray thickener.

veloped primarily as a means of controlling the density of the underflow from the Dorr thickener in the first continuous treatment of cyanide pulps. It has been proved the most satisfactory method of accomplishing this because of its positive, uniform displacement, which can be regulated at any rate of flow within the limits of its capacity. When the underflow from the thickener is too thin, a simple regulation of the pump will decrease the rate of discharge until its density is correct. Once the pump has been regulated, it is seldom necessary to change the adjustment unless there is a decided change in the feed to the thickener. Cord diaphragms constructed on the same principle as cord tires give appreciably longer life than the original fabric construction. Diaphragm pumps of different types are available from several manufacturers. The diaphragm pump has on various occasions been challenged by the centrifugal pump, but the characteristics of the former as mentioned above show advantages in

101

SLIME TREATMENT

its favor. When a thickener becomes overloaded, the centrifugal pump tends to "pack up" and remove less pulp; the diaphragm pump increases its delivery in terms of solids and so tends to compensate the overload condition. On the other hand, the centrifugal increases its output as pulp density decreases and this is obviously undesirable.'

Describing the history of the diaphragm pump, Luther B. Eames, wellknown engineer and an early associate, writes: To our knowledge the earliest use of the diaphragm pump in metallurgical plants was for recirculating pulp in the loading tank of a Moore filter at the Lundberg, Dorr, and Wilson mill at Terry, S.D. This was of the type sometimes called a pitcher pump. It had flap-type valves which, however, caused irregular operation due to the presence of wood chips and tramp oversize in the pulps discharged from thickeners. As a result, pumps were designed more suited to use in connection with thickeners. At the Hollinger mill valves were developed which discharged around the whole periphery of the valve seat and were so designed that the center of gravity of the valve was below the valve seat. Also seats and valve disks were of soft rubber so that any chips or tramp would not cause leakage and would be washed off the seat at the next stroke. This is of importance, particularly in countercurrent washing, where it is essential to keep the capacity of all pumps constant and equal. Ball-type valves have also been used. The balls are of rubber weighted with steel cores and operate against circular rubber seats. This type, however, has been used more generally in the industrial than the metallurgical field. FACTORS AFFECTING RATES OF SETTLEMENT

As discussed in Chap. II, a number of chemical and physical factors affect the settling" rate of ore particles suspended in water or cyanide solution. The use of lime and caustic starch has been mentioned in Chap. II. lt was found at Noranda" that aeration and the presence of sulphates aided pulp settlement: Laboratory settling tests on the feed to the decantation thickeners have shown that sulphates-whether added in solid form, as, for example, (NIlIl 2SO,J, or present in the barren solution, as CaS04-increase the free settling rate about 25 per cent over that obtained with water made alkaline with lime. Laboratory tests and plant operation have also established that a well-aerated pulp settles better than one in which aeration in incomplete. This probably is the reason why thickening rates in the cyanide plant are sometimes appreciably reduced when an ore is being milled which has partially oxidized in the mine and consepuently contains appreciable amounts of reducing salts.

The use of a comparatively new flocculating agent C.M.C. Carboxel, or sodium carboxymethylcellulose, is described by E. E. Brown," mill superin1

A. Clemes, "Modern Metallurgical Practice on the Witwatersrand,"

JOllT.

C.M.

and M.S.S.A., August, 1947. "Cyaniding at N oranda," Trans. 49, C.I.M. and M., 115, 1946. "Sodium Carboxymethyl-Cellulose as a Flocculating Agent for Cyanide Slime Pulps,"Blll. C.I.M. and M., April, 1949, p. 166. 2

3

102


tendent of the Chesterville Mines, Ltd. A water-soluble "ILV." grade used in the amount of 0.01 lb. pel' ton of ore increased the settling rate in the thickeners by 8 to 10 pel' cent. AGITATION

Theory. The mixing of solids suspended in a fluid medium is still largely an art. The development of fundamental laws governing the operation is complicated by the large number of variables involved, some of which can hardly be evaluated in mathematical terms. Practical studies of the circulation patterns of the more common types of agitators are to be found in chemical engineering literature. In one of the more recent articles' on this subject the author summarizes a few rules for agitator design as follows: For ratio of tank diameter divided by impeller diameter, take 4: 1 for simple blending of light material; 3: 1 for the average job, and 2: 1 for heavier density or high viscosity material. The approximate impeller speed should be 700 peripheral feet per minute for turbine-type impellers and 1000 to 1500 ft. per min. for the propeller type. The horsepower requirements can be estimated from data given in the various engineering handbooks. Because in certain cases power imput appears to be directly related to agitator performance, i.e., in gas absorption and emulsification, a rather intensive study has been made of the problem. Thomas Hooker- investigated the power function M versus agitator Reynolds number Re. Secondary dimensions such as pitch, liquid depth, and blade width were expressed on a dimensionless basis by dividing by the primary dimension of the agitator span L. The standard systems chosen for this work were those involving axial flow, tangential flow, and radial flow, and plots showing the above relationship for each system are given, in addition. to a number of the design-factor plots. It is suggested that the agitator power consumption can be predicted for most installations to within ± 20 pel' cent using this method. An earlier paper" reports the results of experimental work on the effect of the number, size, and position of baffles on agitator power consumption. The so-called fully baffled condition, when the input reaches a maximum, is discussed. ·1 Emerson J. Lyons, "Praetica.' Mixer Technology," Chem. Eng. Proq., Vol. 44, No.5, p. 341, May, 1948. 5 "Prediction of Power Consumption for Geometrically Dissimilar Agitators," Chern, Eng. Proq., Vol. 44, No. 11, p. 833, November, 1948. B D. E. Mack and A. E. Kroll, Effect of Baffles on Agitator Power Consumption," Chem, Eng. P1·Og., Vol. 44, No.3, p. 189, March, 1948.

SLIME TREATMENT

103

In cyanidation work, however, it is evident that agitation has two purposes only: (1), to keep the solids in suspension while dissolutionis taking place and (2), to supply the required amount of oxygen, Violent agitation has no recognized value in itself, but ample oxygen supply is essential to promote the reaction between gold, silver, and cyanide, Control of aeration during agitation is essential,since excessive aeration usually results in greater consumption of lime and sometimes of cyanide, with no increase in extraction., Length of contact and dilution during agitation vary considerably for different ores. For gold ores agitation periods vary from 16 to 48 hr., probably averaging 30 hr., and for silver ores/nearly twice this period. Dilution during agitation varies from 1 of solution to 1 of solids by weight to 3 of solution to 1 of solids, occasionally higher, probably averaging 2 to 1. Operation is continuous through a series of agitators, preferably three or more, to minimize any tendency for short-circuiting. Change of solution during agitation by means of a thickener placed between agitators often improves extraction, particularly when silver or other ores requiring long periods of agitation are being treated. Three types of agitators are now in general use: mechanical types (Devereux, Turbo, Wallace); air-lift types (Pachuca, Parrnl) ; airmechanical type (Don, and Notanda type). Mechanical Agitators. The first agitators used in the industry were of the simple rotating-paddle type. These, however, required a large amount of power and did not give uniform suspension or satisfactory means for controlling aeration in cyanidation. Devereux. The Deverell.'{ is a propeller-type agitator in which a propeller revolving centrally in a round, flat-bottomed tank at about one-third the depth of the tank above the bottom forces the pulp downward to the bottom where it is deflected along the bottom and up the tank sides, creating a vortex at the center in which air is drawn into the pulp. This machine, while simple mechanically, requires relatively high power. At one plant an 18- by I8-ft. Devereux requires 15 hp. when agitating a pulp containing 62 per cent solids. Turbo and Wallace. The Turbo and the Wallace are high-speed impeller types of agitator imparting high velocity to the pulp and are suited for relatively small-diameter flat-bottomed tanks. The Turbo may be provided with an impeller near the pulp surface whereby large amounts of free air may be drawn into the pulp and emulsified. The Wallace consists essentially of an impeller or rotor near the tank bottom and at the lower end of a stationary vertical pipe which extends nearly to the pulp surface. The action of the impeller creates a circulation of pulp down the pipe and into the impeller. When the pipe is

104


properly placed, free air is drawn into the top of the pipe together with pulp and disseminated throughout the tank. The pipe may have auxiliary openings to draw in pulp from different levels in the tank. Both of these agitators have been used in combination with DOlT agitators for supplying large quantities of air to the pulp when treating ores or concentrate requiring an unusual degree of aeration. Pachuca. The Brown or Pachuca agitator, as it is most commonly known, originated in New Zealand in the early days of cyanidation, the invention of F. C. Brown. It later became popular in Mexico, and to a limited extent its use has spread to other countries. The Pachuca agitator is an all-air agitator which consists of a central stationary air-lift pipe in a circular cone-bottomed tank of varying dimensions, usually from 5 to 18 ft. in diameter and 20 to 60 ft. in depth. TABLE

Size, ft.

15 15 18 15

by by by by

45 60 22 40

12.

PACHUCA AGITATORS-Am-coNsuMPTION DATA

Cu. ft. air per min.

65 122 79 . 70

Air Pressure

.... 42.5 38 14

Dilution

0.6:1 1.2:1 1.4:1 ... , .

Cu. ft. air per 1000-cu.-ft. tank

7.0 11.0 11.6 9.0

Relatively large quantities of high-pressure air are required, "and with some ores considerable difficulty results from the building up and sloughing off of solid masses from the tank sides. The height of the Paehuca isa further disadvantage and usually necessitates pumping of the pulp. The Pachuca agitator is still to a large extent retained on the Rand, about one-half of the continuous-treatment slime plants using it in their flow sheets. A recent paper? states "where maximum aeration of pulp is essential for optimum extraction, it appears that the Brown tank (Pachuca) is superior (to the Don) in maintaining the oxygen content of the pulp at a higher level." This statement is based on some recent work at the the Geduld Propietory mines and the Sub Nigel, Ltd., but tests made in the early 1920's indicated that, if additional aeration is supplied to the submerged rake arms, the DOlT agitator can be made as effective an aerator as the Pachuca and at a lower operating cost owing to lower total air requirements and the fact that it is not necessary to elevate the pulp. The Dorr Agitator. The Don agitator, in general use today, is a combination mechanical and air agitator designed for use in round, flatbottomed tanks as large as 50 ft. in diameter and 25 ft. in depth. 7 A. Clemes, "Modern Metallurgical Practice on the Witwatersrand," Jour, C.M. and M.B.B.A., August, 1947.

SLIME TREATMENT

105

The pulp and coarse solids from the bottom of the tank are elevated by means of air through the revolving hollow central shaft to a head casting above the pulp level and thence distributed through two revolving launders over the surface of the pulp. Coarse solids settling on the bottom of the tank are raked to the central lift by two plow ai-ms attached to the shaft. The rate of circulation in the tank is controlled by means of the air supply to the central lift.

FIG. 29. Phantom view of type A Don' Agitator.

Air is absorbed in the solution both from the air lift and from the series of pulp streams plunging into the surface of the pulp from the distributing launders. Figure 29 shows the more recent type of DOff agitator with I-beam supports, which requires minimum headroom and makes for ready accessibility. The power required for both mechanical operation and air for agitators from 12 to 40 ft. in diameter will vary from about 1.0 to 0.5 hp. per 1000 cu. ft. of capacity. Noranda Type. This agitator" was developed at N oranda mines to handle the heavy sulphide concentrates which required also a high degree 8

"Cyaniding at Noranda," Trans, 49, e.I.M. and lIf., 104, 1946.

CYANIDATION AND CONCENTRAUON OF ORES

of aeration for optimum extraction and economy in reagents. In this machine the settled sands raked to the center of the relatively deep tank are returned to the top of the agitator by an outside air lift while additional aeration is supplied by a series of jets placed around the tank and connected to a circular header above. Tank areas and total air supply are given in Table 13. Air is supplied to the agitators at 10 lb. per sq. in. This pressure was selected because it appears to give satisfactory oxygenation at relatively low cost. 13.

TABLE

No. of tanks

Diameter

OPERATING

DATA-N aRANDA-TYPE

Cireul.

Radial

Total

air

air pipes

urea, sq. ft.

lifts

Total air, ell. ft.

*

AGITATOR

Motor

per min.]

Rake, r.p.m.

--- ---

6 9 12 18

I

3 2 1

* Total t Total

ft. ft. ft. ft.

area air

6 3 0 0

in. in. in. in.

1 2 2 3

6 8 8 12

33.2 201.6 226.2 254.5

93 564 633 713

1 1 1 2

hp., hp., hp., hp.,

500 r.p.m 500r.p.m 500 r.p.m 500 r.p.m

6 6 6 5

R2 of tanks. total area X 2.8.

1r

CONTINUOUS VERSUS BATCH AGITATION

A common problem in mill design is to determine from the experimental data obtained by batch tests the number and capacity of agitators that must be used in a continuous system in order to approximate the same degree of dissolution. A mathematical analysis of this problem is described by McMullin and Weber in Trans. 31, A.I.Ch.E., 409, 1935. As a result of further studies along these lines, Dr. E. J. Roberts" has prepared the graph shown in Fig. 30, which expresses the relationship between the ratio of volumes for continuous and batch systems and the percentage of incompletion of the reaction for series of one to four agitators. Commenting on this relationship, Dr. Roberts points out that the numerical values obtained should be taken for general guidance only because the assumptions involved in the calculations are first approximations only based on the hypothesis that the rate of dissolution is a function of the continuously diminishing surface of the substance being dissolved. In the case of the extraction of gold by cyanide solution, it is probable that other factors such as the rate of diffusion through pores in the rock particles, intensity of agitation, etc., affect the over-all rate of 9

Research director, Westport mill laboratories.

107

SLIME TREATMENT

dissolution. For precise design, therefore, it is recommended that the MacMullin method be followed, which involves running tests to determine the type of reaction empirically. As an example of how Fig. 30 may be used, take the case of -gold being dissolved by cyanide solution. Assume 98 per cent completion of reaction is desired, or 2 per cent incompletion. Referring to the graph, the required reaction could be completed in one agitator if the volume relationship was 6 to 1. In other words, if 10 hr. was required by batch agitation to extract 98 per cent of the gold which could be extracted in infinite time, a single continuous agitator would have to be large enough

.,

60 ~

.a

40

30 ~ 20

s~

~

§

III

15 .C)

~

<,

r-,

No. of agitators ........ in series

'1'-

.

/0 8 6

§ ~

~

<,

",,/

......: f.-.

3

2 / .07 QI

r-- I t--- I--

........

r-, I"

2

1.,. r--

I"--...

f.-..

r-

j .2.3 .4 .5.6 .8 /

2

I'--. t::-- t-- r--...

3 4 56 8 /0

20 3040

Percent incompletion

FIG. 30. Chart showing relationship between ratios of volumes required for batch and continuous agitation (in the case of one or more agitators in series) for various percentages of gold extraction.

to give 60-hr. detention time. However, if two agitators were used in series, the combined volumes would have to be large enough to give only 21.6-hr. detention, while in the case of three agitators in series the time is reduced to 16.4 hr., or 5.5 hr. for each agitator. In actual practice, an economic balance covering first cost and operating costs for the various combinations of agitators indicated would have to be worked out and the most favorable of these selected. FILTRATION

Definition. Filtration may be defined as the separation of liquids from solids by passing the liquid vehicle through a porous medium, which offers free passage to the liquid but refuses passage to the solids suspended

108


therein. Either the solids or the liquid constituent or frequently both may be products of value, and the ratio of liquid to suspended solids in the feed pulp may vary between extremely wide limits. In metallurgical work, however, the filter is usually preceded by a thickener to dewater the pulp as far as possible by gravity, thereby producing a uniform feed and one on "which the filter can operate at its greatest capacity and give best washing results. Principles. (1) It is evident that a differential in pressure must exist on the two sides of the filter medium in order that the liquid will flow through the pores of the filter medium and the retained solids. (2) Once this layer of solids has formed, its surface acts as the filter medium, and initially cloudy or dirty filtrates, due to the passage of solids through the relatively large openings of the filter medium, tend to clear up as the solids bridge these openings. (3) Since the minute voids between the solid particles in the filter cake resemble capillaries, the flow of liquid may be expressed by a modified form of Poiseuille's equation." The instantaneous rate of filtration per unit area can be regarded as the ratio of the pressure to the product of viscosity by the sum of cake resistance and filter resistance. (4) For incompressible cakes, the rate of flow is directly proportional to the area and inversely to the viscosity and cake thickness. For flocculent or slimy materials, however, the rate may increase but slightly with pressure increase and in some cases actually decrease above a certain critical pressure. (5) In general, the thinner the cake the higher the rate of flow, but other considerations, such as washing and drying efficiency and ease of discharge, usually call for a cake of some optimum thickness determined by test. (6) The rate of cake formation is also proportional to the ratio of solids to liquid in the feed, but filter capacity in terms of either solids filtered or filtrate clarified is dependent upon completion of the entire cycle, including cake formation, washing, drying, and discharging. Filtration Media. These may be divided into two main classes-the fabric and the metallic media. Of the former, the most commonly used are of cotton, different weights of duck, twill cloth, and special weaves, sometimes chemically treated for special purposes. Synthetic-fiber weaves are now widely used. Some, such as vinyon, are resistant to acid attack. Vinyon-N resists considerable heat as well. Rayon, Nylon, and Saran are excellent fibers for certain uses, and all these synthetic materials are available in a wide range of weaves. Al10 This equation can be integrated for constant pressure filtration, or for constant rate' of filtrate flow. For the mathematical development see "Theory of Filtration" by Hugh Bellas, in John H. Perry, Chemical Engineers' Htuulbook, 3d ed., McGrawHill, 1950.

SLIME TREATMENT

109

though they are more expensive than the natural fibers; increased life . often more than compensates for this greater first cost. Of the metals, there are woven-stainless-steel and monel-metal screens wdw~. . The cotton ducks are satisfactory and are used mainly on pressure-type filters. The twill cloths are resistant to abrasion and are widely used on vacuum filters. The special weaves-chain, square, and basketfind their chief application on vacuum filters where high capacities must be obtained with coarse solids and a clear filtrate is not essential. Factors governing the selection of a filtering medium are J 1. It must have mechanical strength and resistance to the solution to be filtered and have a satisfactory economic life. This determines the weight and type of the material. 2. It must produce a filtrate of desired clarity and retain the smallest particle in the feed. This determines the grade and weave of the cloth. Filter Cloth. The texture of filter cloth as used in filters in cyanidation for some years has been largely that of so-called twills, and in these cotton fabrics the weight usually runs from 15 to 1772 oz. The texture as such would be difficult to describe except that it produces a diagonal ribbing of characteristic appearance with which all cyanide operators are familiar. This is produced by alternately skipping certain threads in the weaving . which on the next threads are not omitted. At an earlier period the so-called square-weave duck fabrics were used, but they are distinctly unsuited for this work, being better fitted to retain liquids than to permit their passage. In more recent years there has been a tendency toward lighter material and the synthetics, and with better general understanding of filter fabrics this tendency may proceed still further. Meanwhile, it remains a matter for determination by trial as to the best fabric in a given case. The price has fluctuated with conditions in the cotton-textile industry. Generally, the material is sold in make-up lengths as well as being stitched for a given width of filter in the case of the Oliver machine. Nowadays, seamless filter cloth up to 14 and 16 ft. in width can be supplied. Occasionally, material is sold in rolls for the mining company to make up its own covers, but the customary practice is to buy the cover readymade. In the case of the American disk filters, which use a specially tailored bag to cover the individual sectors, the covers are made up to fit and are sold by the dozen or other convenient multiple. The bags used on the American filters are occasionally heavier than the covers used on the Oliver filters because at times the wear is greater. As to the useful period for cloth on cyanide filters, there is a wide variation; a filter life as long as 220 to 230 days has been reported in the case

no


of one large installation-a, remarkable performance. From this figure of nearly 8 months we go down through various periods, finding that 150 to 160 days is considered excellent life and that a good many plants are obtaining 100 to 120 days. Where working conditions are severe, the life is less than 100 days, but possibly in all such cases means might be found to increase it. In removing lime encrustations by hydrochloric acid, 5 per cent is considered too concentrated for safety, and 2 to 27~ per cent is preferable, with reasonably frequent treatments, instead of stronger acids, being used at longer intervals. The frequency of acid treatment is a factor that varies considerably in different places. In northern Ontario it is used much more often than in other places where lime concentration is less. In this respect it resembles the frequency of acid treatment necessary in the silver cyanide plants in southern Mexico where extremely high lime concentration prevails. Stage Filtration. Two and sometimes three stages of vacuum filters have been used in series with pulping apparatus between each stage for washing pulp. Diluting liquid is added to each filter cake ahead of each repulper. In some cases partial C.C.D. is employed prior to the filtration stage to reduce the value or grade of liquid going to the filters. Sometimes there are short periods of agitation between the filters. This method of working has been adopted by several of the Canadian cyanide mills, but its use has not been general elsewhere. Operating costs are relatively high compared with C.C.D. or C.C.D. followed by a single dewatering filter. For examples see descriptions of the Lake Shore, WrightHargreaves, and Hollinger cyanide mills in Chap. XV. Filter Types. Two types of filters are in use-those which employ pressure and operate intermittently, with definite cycles for charging, washing, and discharging; others that employ vacuum and operate continuously or intermittently, with synchronized feeding, cake forming, washing, and discharging. Among the filter presses are the Dehne, Shriver, Merrill, Kelly, Sweetland, and Burt. In the metallurgical field the Dehne found some use in Australia and the Merrill is still in limited use in North America and Central America. Among the vacuum filters in use today are the Moore and Butters leaf filter of the intermittent or cyclic type, the continuous revolving types, including the external drum filters made by Oliver and a number of other manufacturers; the disk type or American filter; the internal drum or Dorrco filter; and finally the more recent horizontal type in which filter pallets revolve around a vertical axis. Pressure Filters. Dehne Press. The Dehne plate-and-frame filter press has satisfactorily filtered many millions of tons of raw and roasted slime in Australia, particularly in the western states. At Kalgoorlie 100 of them

I,

SLIME TREATMENT

111

were at one time treating 100,000 tons a month. Their operation was described in detail by M. W. von Bernewitz in 1906 in Proc, Australian l.ll1.E., now the A .1.111. tuul lJ1. Most of these presses were of 5-ton capacity per charge. A few were hydraulically closed, but most of them were hand closed. In their discharge much hand labor was used. They can be emptied slowly by reversing the washing valves and opening the filling valve. Pressure filters are expensive in operation, and good results are dependent on careful attention and manipulation on the part of the operator.. They have been almost entirely replaced by the continuous vacuum type. Merrill Press. The Merrill filter press (see Fig. 31) is essentially a plate-and-frame press but is practically automatic in filling and discharge. Unlike the Dehne, the Merrill does not need opening save for renewing cloth or making other minor repairs. Along the median line at the bottom of the press, passing through each plate and frame, is a continuous channel within which is a sluice pipe bearing a series of nozzles, one for each frame. After a cake has been formed and washed, the sluice pipe is oscillated by a small motor through an arc of 180 deg., and water under pressure is discharged from the nozzles. The slime thus washed out of the frame is repulped with the water and discharged from the press through a number of cocks leading from the annular sluicing chamber. Thirty-one Merrill presses at the Homestake mine, South Dakota, have the capacity to treat-not merely wash and filter-at least 1500 tons a day of slime of which 99 per cent passes 200 mesh. Each of these presses has 90 frames, 6 by 4 ft. by 4 in., the last dimension being the thickness of the slime cake. They hold 26 tons of dry slime. The slime is dewatered at the stamp mills to 35 per cent solids and flows by gravity-to the slimetreatment plant three miles distant. The press plates are covered by two cloths-a light muslin twill next to the frame and No. 10 cotton duck which covers the muslin; the muslin prevents the duck from being forced into the plate corrugations. The cotton has a life of 20 months. A typical treatment cycle is shown in Table 14. While a few other plants use Merrill presses for washing, the Homestake is the only plant practicing direct treatment in Merrill presses. VACUUM FILTERS

Leaf Type. Moore and Butters. The Moore was the first of the leaf filters and was developed in 1901. The individual leaf consists of a frame formed of a perforated pipe through which suction and compressed air are applied. The pipe is covered with a filtering medium both sides of which are sewed together in equally spaced vertical seams. Wooden slats are inserted into the pockets between these seams to prevent collapse of the

E. F. G.

D.

C.

A. B.

i~?:

--=

~r{"J~nff~;~m;:';;';!;'Gf":ttr;-w~~r$~.i')'

FIG. 31. Merrill automatic sluicing-pressure slime filter-part filling or center-washing type. Standard container or frame. H. Sluicing pipe, containing water under 60- to 90-lb. presFeed channel through which the slime pulp enters each sure admitted at either or both ends. frame. T. Sluicing motor and gearing. Pipe from which water or solution is drawn off during J. Residue discharge cocks. process of filtration. K. Effluent solution launders. Slime or pulp line. L. Discharged residue launder. Barren-solution wash line. M. Peep gate. Filter plate. N. Lever system for residue discharge cocks. Water wash line.

;'rt}tM~~~~:~~'ii

D

~

[f1

toJ

o;
o

bj

Z

o

~

toJ Z

a o Z a

~tj

Z

o

~

tj

Z H

a

I>:l

,..... ,.....

SLIME TREATMENT

14.

TABLE

113

SLIME-TREATMENT CYCLE AT THE HOMESTAKE

Period Operation

Remarks Hours

Minutes

.

1

o

First aerating ..... , .....

1

30

Filling

Strong-solution leaching

.

30

Second aerating

.

30

Second strong-solution leaching .

Third aerating

_

.

Third low-solution leaching

.

Washing

.

Sluicing Total cycle

. .

2

0

30

3

1 10

Discharge from sludge pressure tank to presses contains 35% solids. Gage pressure, 26 lb. Effluent partly to waste and partly to storage for wash ing. Gage pressure, 25 to 30 lb. Effluent to waste "and to storage. Lixivium: 0.055 % NaCN; effluent to low-solution pumps, followed by clari fication and precipitation. Barren solution used as preliminary wash. Gage pressure, 25 lb. Effluent to weak solution sumps followed by clarification and precipitation. Barren solution Ore used. Gage pressure, 30 lb. Effluent to weak solution sumps, followed by clarifi cation and precipitation. Barren solu tion reused. Gage pressure, 30 lb. Effluent to weak solution sumps, followed by clarifica tion and precipitation. Barren solu tion reused.

0

Lixivium: 0.035 to 0.04% N aCN; gage pressure, 30 lb. Effluent to weak solution sumps, followed by clarifica tion and precipitation. Barren solu tion reused. 15 to 20 Solution displaced during press filling is raised to full cyanide strength in strong solution sump, then to storage tanks Wash effluent to strong-solution make up. Amount varied to balanco solu tion needs. 30 Residue pulp to discharge. 50

two sides under vacuum and to assist drainage. An assembly of such leaves is known as a "basket." Four tanks having vertical sides superimposed on tapering bottoms are required. The first tank holds the slime

114


to be filtered, the second is filled with barren solution for displacement of the gold-bearing solution in the cake, the third tank is filled with water for displacing the barren solution, and the fourth is used for receiving the washed slime which is loosened from the leaves by compressed ail'. This residue may be trammed or pumped to the dump. The basket is raised and lowered and moved from tank to tank by means of a chain hoist mounted on an overhead runway extending the full length of the four tanks. The Butters filter has been used to a greater extent than the Moore, but both have now been largely replaced by C.C.D. and continuous filters such as the Oliver drum type. The Butters filter is similar in construction to the Moore, but the basket always remains in one tank. Into and from this, according to the cycle, are pumped and discharged slime and washing solutions or water. The treated slime is discharged as in the Moore system. The Butters filter is still used in the older plants in South Africa. A typical example is found at West Springs mines, where an average of 3685 tons pel' day is treated, the ore being ground to 66 per cent minus 200 mesh. After thickening and agitation the pulp is filtered on 490 Butters leaves. A total of 4050 tons per day of barren solution is used. The filter discharge contains 30 per cent moisture and assays 0.02 dwt. dissolved gold and 0.17 dwt. undissolved gold pel' dry ton. Capacity of the filter is about 85 lb. dry solids per day pel' square foot of canvas. Rotary Type. American Filter. This machine consists of a number of parallel, individual, cloth-covered disks mounted on a hollow shaft through which suction and compressed ail' may be applied. The lower halves of the disks are submerged in the slime tank. The disks are divided into independent and removable sectors. Wash water may be applied from spray nozzles. The assembled disk and shaft section is shown in Fig. 32 (see also Fig. 33). The American filter is a continuous machine of mechanical simplicity. It occupies small floor space and presents a larger filter area or surface in proportion to space filled than any other filter. Inflation of the filter bags during each discharge period keeps the cloth in good condition, and the cake discharged has low moisture content. The changing of filter cloths and the sectors is done easily and quickly. No pulp agitator is needed in the tank. This filter is made by the Oliver United Filters in four sizes: 4, 6, 8~, and 12~ ft. in diameter with 1 to 12 disks. In operation the thickness of the cake ranges from 7~ to % in., depending on the material being filtered. As the disk slowly revolves, each sector in sequence rises from the pulp with vacuum still applied. Fine sprays of solution or water then are applied to the cake, thereby displacing the original liquid left in the pores of the cake. After a sector passes

115

SLIME TREATMENT

CL,.A".4P

BAG

CLAM P --t

~:;;~B;-"';"---~~-2'-_

BOLT CIRCt..£

vr'l~=======:il-CENTER SHAFT

...

SECTION

__-I-_ FILTRATE

~--_---'---'=";"'

CHANN~

FIG. 32. Assembled disk and shaft section Of American filter.

FIG. 33. An early installation of American filters

III

Canada.

116


~----13 ': I it/O. overoll-----l~

14~OV; overall

Roll guards furnished

Pulsating valve 2-IY2" std female pipe conns. connect one to law pressure air,-one to vacuum line.

Nate, See foundationplan fqr location of drive sprocket. Locate drive unit at convenient level; Chain furnished for 5'-0" ctrs; see separate drawing fo.r drive unil.

VALVE END

8~6"O. (OPflning in drum)

12'-5~"o. inside Overflow conn. :;1"std flgd conn. 8-~-1I Me. tapped hales an

Shell head wipers Overflow launder

7y:!" B.e. holes straddle

t{s.

DISCHARGE END FIG.

34. The Dorr filter-side and

SLIME TREATMENT

117

Overflow lip Discharge chute lip

To overflow conn.

SIDE ELEVATION

Valve end

-'%.

4 bolts 2~4'projer;tion allowing for ~4·grout (Customer Iornlsbed]

FOUNDATION PLAN end elevations and foundation plan.

118


out of the spray zone, vacuum still applied, a large portion of the wash that has replaced the original liquid is also drawn out. Now as dryas it can be made, the cake is in propel' condition for discharging. This is done by cutting off the vacuum and introducing a small quantity of 2 to 3 oz. air, which slightly inflates the bag. The cake will be so loose that it can either drop off or be scraped or rolled off easily. In the metallurgical field the American filter finds its principal use today in dewatering of flotation concentrates. Dorrco Filter. This machine is of the rotary-drum type, but the filtering medium is applied on the interior of the drum which acts as its own container. The inner surface of the drum is di vided into a series of fabriccovered panels, each forming a face of a polygon inscribed wi thin the cylinder. The compartments thus formed connect individually with an automatic valve, which, in turn, applies vacuum and compressed air in proper sequence. The closed end of the drum is carried by a trunnion and bearing, and the open end is equipped with a steel tire which runs on rollers. An annular ring forms an internal flange at the open end, serving as a dam for retaining the pulp being filtered. Feed may be introduced by a pipe or by a trough passing through the open end. Cake is discharged by compressed air into an internal hopper, equipped for either chute discharge or discharge on to a conveyor belt or screw conveyor. Side and end elevation drawings of the Dorrco filter are shown in Fig. 34. Oliver United Filters makes these machines in six diameters, from 4 to 14 ft., and up to 18 ft. in length. At the Pioneer mill, British Columbia, the practice in maintaining the Dorrco filter was to treat the filter cloth-a No. 26 twill-with acid every 21 days. The procedure consisted of hosing out the filter, dumping in a dilute solution of hydrochloric acid, and rotating the filter drum for an hour; the filter was then drained, again hosed out, and put into service. There was no scrubbing of the cloth, the pores being kept open easily throughout its life. To keep the cloth in best condition, low-pressure air was blown througb it for several seconds after the cake had been discharged. The filter cloth was changed after 108 days of service, two men making the change in about 8 hr. Each filter had 371 sq. ft. of canvas area, which, on the basis of washing 200 tons of solids per day, showed a capacity of 1078 lb. per sq. ft. canvas pel' 24 hr. Oliver Filter. This machine consists of a cylindrical drum divided at its periphery into a number of individual compartments and covered with a cotton material, held in place by a winding of wire. The lower portion of the drum is immersed in a tank filled with thickened pulp to be filtered. The pulp is fed to the tanks continuously and is maintained in suspension by reciprocating stirrer bars. The drum is supported by two trunnion

119

SLIME 'l'REATMENT

bearings and is revolved slowly through a worm and gear drive. The interior of each compartment of the filter shell communicates'through a separate conduit to a rotary port valve mounted at one end of the drum. Through this valve either suction or positive air pressure is applied to the different compartments in the propel' sequence and for the desired period. The slime cake is discharged from the chum by a stationary scraper blade which barely touches the cloth at the point of discharge. Compressed air is admitted just ahead of this position to loosen the cake. Sprays or drip pipes for washing the cake may be applied through nozzles mounted on the frame. A wash net riding on the surface of the cake and upon which oJ

FIG. 35. An installation of Oliver filters at the Hollinger mill, Ontario, Canada.

the drip falls serves to distribute the water evenly over the cake. Figure 35 shows some Oliver filters at work. The Oliver-type or drum filter is now more generally used than any other type in cyanide plants for washing and final dewatering. At Wright-Hargreaves in order to prevent high dissolved loss, the filters must be scrubbed and acid-treated about every second day. The scrubbing was formerly done by two men and required more than 2 hr. to make a good" job of scrubbing and acid-treating. A machine was' developed on the property to do this work, and with its use one man can now do the job in about 30 min. Before scrubbing is started, the filter is reversed, and the scrubbing machine laid against the back at an angle of about 5 deg. As the scrubbing proceeds, weak hydrochloric acid is sprayed ahead of the brushes. The average life of the covers is a little more than 2 months.

120


Use of Flappers. At Lake Shore, the concentrate filtration is improved by the use of a device described by J. E. Williamsen!' as follows: The concentrate is filtered on two 8 by 10 ft. Oliver filters. One of these filters is used for each roaster. The filter cake discharges onto a conveyor belt which feeds the charge directly to the roaster. The concentrate filter cake is at best a sticky, puttylike mass. If the moisture exceeds 20 per cent it becomes difficult to handle the cake on the belt and at the charging chute. The best filtering technique which could be obtained on the 26 per cent sulphur concentrate produced a cake with from 21 to 22 per cent moisture. This was too sticky for continuous use and introduced an excessive amount of moisture into the furnace, which slowed the commencement of roa sting. To reduce the moisture content of the cake, a device was copied from the Consolidated Mining and Smelting Company in their Sullivan concentrator, which, according to them, was originally developed at Granby. This device consists of a flapper fixed on a revolving shaft parallel to the axis of the filter drum. The flappers are made of old conveyor belt about 12 in. wide and the length of the filter. The single flapper reduces the moisture from 21 to 22 per cent down to 16 to 17 per cent, at which moisture content the concentrate can be handled and charged with a minimum of difficulty. The flapper is run in the opposite direction of rotation to the drum. When both were run clockwise, the blow from the flapper tended to stop the drum, making for jerky operation. The flapper is driven by a I-hp. motor which is probably more than twice as large as is required, as the same motor was sufficient when two flappers were fitted on the filters.

To prevent the cake from being dislodged by the blows of the flapper, a piece of heavy fabric is usually fastened so that it drags upon the cake as the drum revolves and receives the blows given by the flapper. The blows cause a rearrangement of the particles in the cake, closing the cracks and liberating moisture, which is then drawn through into the vacuum system. Compression Belts and Rolls. Compression rolls, with or without an endless-belt attachment, mounted on a frame above a drum filter so that they exert pressure on the filter cake are used to close cracks in the cake and reduce the moisture content. A still more recent development for reducing the moisture in the cake on a rotary-drum filter is the use of a vibrating shoe which rests on the surface of the cake. Flood Washing. Flood washing the residue filter cakes before discharge using an excess of water for lowering the dissolved gold losses by displacement is in use at Wright-Hargreaves and a number of other Cana11 "Roasting and Flotation Practice in the Lake Shore Mines Sulphide Treatment Plant," C.I.M.M. meeting, January, 1947.

121

SLIME TREATMENT

dian plants. A so-called excess wash s'coop, or launder, extending the full width of the filter is placed just above the tank level on the ascending side of the drum. The water from the sprays, which are mounted just below the highest point of the drum, floods down over the surface of the cake and is deflected into the launder by means of a rubber lip, thus preventing dilution of the pulp in the tank. The excess water caught in this manner 'is then returned to the mill or wash-water circuit. TABLE

Plant

15.

RAND FILTER PRACTICE

A

Feed pulp specific gravity ........ 1.434 Filter cake, per cent moisture .... 24.5 Precipitation ratio ............... 1.575 Barren wash ratio ................ 0.91 Assay caking effluent, dwt ........ 5.0 Assay barren solution, dwt ....... 0.015 Assay pregnant solution, dwt ..... 3.1724 Assay tails solution, dwt .......... 0.0520 Dissolved loss per dry ton, dwt... 0.0169 Per cent r cpl acemcnt.] ............ 99.62 Per cent filter recovery ........... 99.69 Thickener discharge specific gravity ............................ 1.68 Barren waste ratio ............... 0.225

B

C·

D

E

1.43 28.0 1.60 0.90 10.0 0.02 6.8580 0.1159 0.045 99.04 99.59

1.53 30.0 1.29 0.90 4.0 0.01 2.3830 0.0466 0.020 99.09 99.39

1.45 24.0 1.49 0.80 1.8 0.01 1.2000 0.0660 0.02 96.89 98.91

1.404 33.1 1.39 0.71 1.93 0.01 1.6500 0.0283 0.014 99.05 99.39

1.588 0.231

1.60 0.0

1.58 0.299

1.64 0.118

* Grinding in cyanide solution. t In figuring per cent replacement the value of barren solution was deducted from the value of the solution in the cake after washing. The filter recovery does not allow for loss through wastage of barren solution, and it should not do so, since barren wastage is a result of grinding in water. It should be noted, however, that in at least one plant no water wash was being used even though grinding was carried out in cyanide solution.

Rand Filter Practice. While some of the older plants are using Butters-type filters, about 75 per cent of the filter tonnage is filtered on drum-type continuous filters, most of which are 14 ft. in diameter by 16-ft. face. Washing is with precipitated solution only, since in most of the mills grinding is in a water circuit and it is therefore necessary to waste barren solution to maintain cyanide solution balance. On the continuous filters the amount of barren wash used will average about 0.9 ton per ton of dry slimes, and dissolved loss is usually reported as less than 0.02 dwt. gold (3~ cents) per dry ton. Filter duty is high, ranging from 1400 to 1800 lb. per sq. ft. of filter area. Table 15 gives details of continuous-filter operation at five of the plants.

122


CONTINUOUS COUNTERCURRENT DECANTATION

Principle. The principle of C.C.D. is simply that, when water or solution is to act upon solids, both are made to pass, in contact, in opposite directions, so that at each end the strongest or most potent portion of either is acting upon the weakest or most exhausted portion of the other. The recovery of dissolved gold and silver from slime pulp in the cyanide process, as first practiced, employed intermittent decantation. Attempts to make this process continuous instead of intermittent were made as far back as 1902, but without success. The invention of the Dorr thickener furnished a means of continuous slime settling on a large scale and made the C.C.D. process possible. Although flow sheets and operation will vary with conditions, the following will give a general idea of the C.C.D. process. After the ore has been reduced to a uniform fineness by wet grinding and classification and the major portion of the solution removed, it is agitated continuously in a series of about three agitators with cyanide solution, where most of the gold and silver is dissolved. The mixture of solids and solution is fed continuously to the first thickener of the C.C.D. series, consisting of from three to five thickeners. It is diluted with the overflow solution from the next thickener in relatively large quantity, a solution that has already been in contact with the pulp in all succeeding thickeners and so is comparatively rich in dissolved gold and silver. The overflow from the first thickener of the C.C.D. series containing the greatest amount of dissolved metal is sent to precipitation or to the grinding and classifying circuit. The underflow is pumped to the next thickener, diluted with the overflow from the third of the series, settled and thus proceeds through each thickener, until discharged from the last as tailings. By each step of the system a large amount of low-grade solution is mixed with a very small amount of higher grade solution so that the value of the solution entering each successive thickener is materially reduced. Barren-cyanide solution from the precipitation apparatus is added to the thickener preceding the last and makes up the bulle of the countercurrent flow. Water is added to the last thickener to make up for the liquid losses in the tailing and to cut down cyanide consumption. A thickener called No. 1 precedes the agitators in the mill; therefore the first thickener in the C.C.D. series is usually referred to as No.2. Uses. 1. To recover practically all of the dissolved values from the finely ground solids, without filtration. The flow sheet must be varied somewhat, depending upon the strength of the cyanide solution to be used and the amount of solution required to be precipitated. 2. To reduce the dissolved value in the pulp going to a filter; so that the

SLIME TREATMENT

123

final dissolved loss will be decreased and the use of a barren-solution wash shortened or omitted. This increases the filtering capacity, tliminishes the cost of operating a filter, and lowers the loss of dissolved value. Application. The application of the C.C.D. process is illustrated by the accompanying typical flow sheets. Figure 36 (type B) shows the most common type of flow sheet in use, where the ore settles readily to 50 per cent moisture or less, and where a strong cyanide solution is not required. The. overflow of either thickener W or thickener X may be used for dilution in the agitators without affecting the distribution of values in the system. By using the overflow of X a lower grade of solution is obtained, which may be beneficial, and the dilution of feed to thickener W is decreased. TYPE

B

Dissolved Value Loss. Conditions assumed: (a) 100 tons of ore per day crushed in cyanide solutions. (b) Discharge from all thickeners with 50 per cent moisture. (c) $10 value dissolved per ton of ore.

50 per cent in mill and 50 per cent in agitators. 400 tons of solution from thickener V precipitated to $0.02. Agitation with a dilution of 2 of solution to 1 of solids. Let V, W, X, Y, and Z represent the value in dollars per ton of solution discharged from the respective thickeners.

(Il) (e) (f) (g)

Equating alit of and into Each Thickener: (1) (2) (3) (4) (5)

100V 100W 100X 100Y 100Z

+

400V = 500W = 500X = 100W = 100Z 100Y

+ 600W + 500X + 500Y + 100Z

Simplifying: (1)

V

=

+

(0.50 X $10 X 100)

+ 100W + (0.50 X $10 X 100) + lOOV + 500Y + lOOX + (400 X 0.02) + 100 tons of water value $0 Solving:

W

(2) W = X

+ 1.00

+ 1.20

(3) X = Y + 0.24 (4) Y = 0.2Z 0.064 (5) 2Z = Y

+

V = $2.51111 W = 1.51111 X = 0.31111 Y = 0.07111 Z = 0.03556

By calculation, the following results are obtained: Assay value of the pregnant solution; i.e., value of V = $2.51111. Assay value of the discharged solution; i.e., value of Z = $0.03556. Loss of dissolved value per ton of ore, $0.03556. Dissolved value saved, 99.64 per cent.

124


Meehan'ieal Loss of Cyanide.

Conditions assumed: (a) Neglect the cyanide consumption throughout the system.

(b) Strength of cyanide per ton of solution, 1.0 lb. (c) Let TT, W, X, Y, and Z represent the strength in pounds of cyanide per ton of

solution discharged from the respective thickeners. Equating old of and into Each Thickener:

(1) (2) (3) (4) (5)

TT = 1.0 100W + 100X + 100Y + 100Z +

600W 500X 500Y 100Z

= = = =

100W 100W 100Z 100Y

+ 100TT + + 500Y + 400TT +

Solving:

TT = 1.0 (2) 6W = 5X + 1 (3) 6X = W + 5Y (4)6Y=Z+X+,± (1)

Z

100X

+ 100 tons of water

Simplifying:

(5) 2Z =

500X

Y

TT = 1.0 W = 0.9109 X = 0.8932 Y = 0.8898 Z = 0.4449

= 0.4449 lb. = mechanical loss of cyanide per ton of are

Type B. 500

Precipitation Type BB.

Precipitation Type D. FIG. 36. Countercurrent decantation, typical flow sheets.

125

SLIME TREATMENT

Where the extraction is obtained slowly, and by experiment it is found that an additional change of solution during agitation is beneficial, the flow sheet shown in Fig. 36 (type BB) may be used. This arrangement allows two changes of solution with a thorough washing of the 'pulp during agitation. The pulp during the final agitation is in contact with fresh solution carrying low values, conditions most favorable for dissolving the refractory material in the residues. Type D flow sheet (Fig. 36) is more generally used in cyanide plants today with or without a change of solution during agitation as in type BB. This type of flow sheet is indicated where solutions stronger than 17~ lb. NaCN per ton are used, where the pulp will not settle to at least 50 per cent solids or where a dewatered filter cake is desirable for tailing-disposal reasons. TYPE

D

Dissolved Value Loss. Conditions assumed: 100 tons ore per day crushed in cyanide solution. Discharge from all thickeners with 50 per cent moisture. $10 value dissolved per ton ore. 50 per cent in mill, and 50 per cent in agitators. 300 tons solution from thickener TV precipitated to $0.02. (f) Agitation with a dilution of 2 of solution to 1 of solids. (g) Displacement efficiency of filter, 60 per cent; i.e., 60 per cent of the value of the solution in the solid cake, which is assumed to contain 33~3 'per cent moisture, or 50 tons of solution to 100 tons of solids, is recovered. The 50Z returned from the filter to the last thickener represents 50 tons of solution removed in loading the filter, which will, of course, still have the value of Z. (h).:[:,et TV, X, Y, Z and F represent the value in dollars per ton of solution discharged from the thickeners and filter, respectively.

(a) (b) (e) (d) (e)

By calculation, the following results are obtained: (1) TV = $3.3439 (2) X = 2.0939 (3) Y = 0.5314 (4) Z = 0.1408 (5) F = 0.0563

To check these figures: Amount precipitated from 300 tons at ($3.3439 - $.02) Amount lost in tailings, 50 tons at $0.0563 = Amount due to neglected decimals Amount dissolved 100 tons at $10 = $1000

=

$ 997.17 2.815 $ 999.985 0,015 $1000.000

126


From the foregoing the following results are deduced: Assay value of the pregnant solution, i,e., value of W = $3.3439 Assay value of the discharged solution, i.e., value of F = $0.0563 Loss of dissolved value per ton of are, 5~loo F = $0.02815 Dissolved value saved, 99.72 per cent Mechanical

L088

of Cyanide.

Conditions assumed: (a) Neglect the cyanide consumption throughout the system.

(b) Strength of cyanide per ton of solution in thickener W, 4.0 lb. (c) Let W, X, Y, Z, and F represent the strength in pounds of cyanide per ton of solution discharged from the thickeners and filter, respectively.

By calculation the following results are obtained ; (1) W = 4.0 (2) X = 3.8124 (3) Y = 3.7655

(4) Z = 3.7537 (5) F = 1.5015 5~OO

F = 0.7507 = mechanical loss of cyanide per ton of are.

Dissolution during Washing. Some additional dissolving of gold and silver takes place during the washing of the ore pulp after agitation with cyanide solution, either in C.C.D. thickener series or in filters. This is generally credited to the change in solution which takes place during the washing operation, whereby solution of lower gold and silver content' and more freshly precipitated is brought into contact with the ore. Further appreciable time of contact is provided in the C.C.D. thickeners. Representative data are not available covering additional dissolution in filters, but this does take place to some extent even on continuous rotary filters, where the time of contact is relatively short.

127

SLIME TREATMENT

TABLE 16. DISSOLUTION IN DECAN'I'ING THICKENERS (C.C.D. cyanide plants) Operation

Hollinger'

United Eastern]

Elko Princet,§

MJgul~

Tons treated per day ....... ......... . ...... Hours of agitation ......................... Change of solution during agitation series .. Head value to plant: Gold (Au) ............................... Silver (Ag) .............................. TotaL ............. '" ................. Dissolved in: Thickener 1, Gold ........................ Silver ....................... TotaL ................................. Thickener 2, Gold ........................ Silver ....................... TotaL ................................. Thickener 3, Gold ........................ Silver....................... TotaL ................................. Thickener 4, Gold ........................ Silver ............... " ...... TotaL .................................

250 12 No

253 62 No

53 72 Yes

100· 24 No

$15.27 10.7 $25.97

$3.10 1.15 $4.25

Total dissolved in C.C.D. series ..........

.

$9.00 .

$20.66 ......

...

$9.00

J

$20.66

$0.08

$0.20

.. - .

......

0.08 0.04 .... 0.04 0.03 '" . 0.03 0.03

0.20 0.07 . .....

0.07 0.04

$0.11 0.08 0.1911 0.09 0.02 0.11 0.05§ 0.05 0.10 No

$0.15 0.03 0.18

.. ..

.... ..

0.03

0.02

...... ......

0.03 0.03 0.01 0.04 0.05

$0.18

$0.33

$0.40

$0.26

......

0.04 0.02

In the foregoing, gold values are based on $20.67 per ounce and silver on $1 per ounce. * From unpublished data, 1916, results in one unit of the C.C.D. series. t From A.I.M.E., August,' 1919. :j: From A.I.M.E., August, 1918. § Third thickener (Elko Prince) followed by dewatering drum filter. II Thickener between second and third agitators. ~ From unpublished data from management, 1913.

CHAPTER IX

Concentration Regardless of subsequent treatment, it is considered best to recover free gold as early and as completely as possible in the flow of p1llp. Gravity concentration in various forms is made use of for this purpose and amalqamaiion used for recovering the gold from the gravity concentrate. Concentration is also extensively practiced where a large part 01' all the gold 01' silver is intimately associated with base-metal sulphides, and the technical and economic considerations involved indicate that such a step in the flow sheet shows advantages over direct cyanidation. Flotation methods, with 01' without some gravity concentration, are usually employed for this purpose.

The concentration of gold and silver ores has two principal objects in view: 1. The recovery of free gold. 2. The collecting of gold or silver values which are associated with sulphide minerals into a relatively small product for subsequent treatment. The latter may involve (a) shipping to a smelter, (b) selective grinding for cyaniding in the mill circuit, (c) cyaniding in a separate circuit with or without roasting. Recovery of Free Gold. Methods for the recovery of free and relatively coarse gold can be classified as A. Direct amalgamation, using 1. Mercury added to mortar boxes or ball mills. 2. Amalgamated copper plates. 3. Tray-type amalgamators (Clark-Todd).

B. Gravity concentration, (followed by amalgamation) using 1. 2. 3. 4. 5.

Riffles or sluices. Hydraulic traps. Unit cells and hydraulic cones. Jigs. Corduroy blankets.

In the case of method A, the gold forms an amalgam, from which it is extracted by the following steps: (1) amalgam collection, (2) cleaning of amalgam, (3) pressing or filtration, (4) distillation or retorting, (5) melt- . ing. 128

CONCENTRATION

129

In the case of processes B, the concentrate must be amalgaI).1ated by special methods which usually include a combination of light grinding and contact with mercury in a grinding "pan" or amalgam barrel. Extraction of the gold then follows along the general lines of the steps outlined above. Because the details of these steps are generally known or can "be found described in Taggart's "Handbook of Mineral Dressing," Sec. 14-10, and other textbooks, brief descriptions only of a few typical amalgamation plants are given in Chap. XI. Metallurgical engineers are generally agreed that free gold should preferably be recovered as soon as it is released by grjnding. There are several good reasons for this: Coarse gold tends readily to segregate in various undesirable parts of the milling circuit, where it is subject to possible loss or theft; it dissolves more slowly in cyanide and may require prohibitively long periods of agitation; the gold may exist in a highly refractory or "rusty" condition and fail to dissolve completely in cyanide solution; if allowed to accumulate in classifier beds, sumps, or launders, its recovery entails considerable work in cleaning up a large bulk of sands at regular intervals. The methods used for recovering free gold are discussed more fully in the chapter on "Bullion Recovery." Gravity methods include hydraulic traps and riffles, corduroys, and jigs. Gold in Sulphides. In the case of gold associated with heavy minerals such as sulphides of the metals, the methods usually employed include gravity concentration or flotation. To determine the most suitable treatment scheme in any particular instance requires weighing the advantages and disadvantages of a number of possible flow sheets. The problem is complicated by the fact that frequently both free gold and gold associated with sulphides occur together in the ore and also that cyanidation either before or after concentration may be a necessary step in the treatment scheme for optimum extraction. While flotation offers the cheaper and more efficient method of concentrating sulphides, it requires supplementary gravity equipment for recuvering coarse gold, if present. Where cyanidation is also used, it involves either two separate circuits (since cyanide is a depressant for sulphides) or special methods of reactivation, as at Lake Shore, following cyanidation. Tabling, on the other hand, has the advantage that it can be incorporated into the cyanide circuit itself and, as used at Hollinger, for instance, accomplishes the double purpose of classifying and concentrating the sulphides, so that only the coarser and more refractory sulphides are removed from the circuit for separate finer regrind and cyanidation, while

130


the finer sulphides are already sufficiently ground for direct cyanide treatment. If, as in many plants today, a final tailing for discard can be made by flotation, there is no question but that this flow sheet, which may include jigs or other gravity equipment for recovering free gold, is the one to use. There are obviously many advantages to be gained in having a small 'bulk of concentrate to grind and cyanide compared with treating the whole tonnage of ore mined. GRAVITY CONCENTRATION

Riffle or Sluice. This somewhat primitive device which has been in use since earliest times still finds application as a gold saver for small prospect or placer mining operations and on gold dredges. The sluice is essentially an inclined trough or launder, the bottom of which is provided with transverse strips, or riffles, with variable spacing. There are endless variations of riffling.' The mixture of sand and gravel is washed through by means of a stream of water, and the heavy minerals, "black sands" etc., along with free gold tend to collect in the spaces between the riffles, while the lighter and coarser material is displaced over the top of the riffles. Periodically the feed is deflected and the heavy concentrate dug out by hand to be cleaned up by anyone of the various methods later described. Hydraulic Gold Trap. The hydraulic trap, another method of catching gold, is described by Ernest Gayford in Trans. 112, A.I.M.E., 1934, and is shown in Fig. 37. It is a simple device inserted in the mill-classifier circuit; sometimes two or more are used in series. At the MontezumaApex plant, Nashville, Calif., the 55-mesh discharge from each of two Marcy mills passed through five hydraulic traps, followed by corduroy tables, Don classifiers, and flotation. The traps were bled daily and thoroughly cleaned out twice a month. The product was concentrated periodically on a small Wilfley table. The concentrate was then ground and amalgamated in a barrel, and the tailing returned to the mill circuit. Cones. At the cyanide plant of the Pickle Crow gold mine' in northern Ontario three cones are built into the launder between the ball mill and classifier. Each cone is 18 in. in diameter with a slope of 50 deg. and fitted at the apex with a 4-in. nipple and discharge valve. A solution supply is connected into this nipple, and a regulated flow keeps the contents of the cone free of slime. The cones are dumped every 8 hr. under normal operating conditions, and the concentrate obtained varies from a few ounces to as high as 200 oz. of gold pel' ton. The Unit Flotation Cell and Hydraulic Cone. To save coarse gold and recover free minerals at an early stage in milling, some plants installed 1

See Taggart, Handbook of Mineral Dressing, Sec. 11-95, Wiley, 1945.

131

CONCENTRATION

the Denver Equipment Sub-A flotation cell and hydraulic cone. Figure 38 shows its place in the circuit at the McIntyre-Porcupine mine, 'Ontario. The flotation machine and its use in the circuit have been patented in several countries.

Locking (hinge

f

• 2f" warer line

"".,....-aJ-I--H-----T_ _ .JJ

4"/ubricOiting --~~~1-- -1---1~(;(Ive locking type

FIG. 37. Hydraulic gold trap.

The unit flotation cell prevents the accumulation of free gold in a ballmill classifier circuit by removing coarse gold as soon as possible. Otherwise, because of its high specific gravity, the gold remains in this circuit until finely ground. Also, periodic surges of the classifier may allow this accumulation of gold to get into the mill circuit with resulting loss because Coarse free gold is not readily floated and is often slowly extracted by cyanide. Moreover, the gold sufficiently fine to overflow the classifier lip

132


is likely to become coated with sulphide slime during the grinding operation, reducing the recovery by flotation. Jigs. During the last twenty years the use of a jig in the mill-classifier circuit to remove coarse gold has become a very general practice. In addition, jigs ani widely used in placer operations to replace sluice treatment in the recovery of fine gold. Modern jig design has much.improved the metallurgical and mechanical efficiency of these machines. In the mill circuit, the jig offers the advantages of very small floor space requirements and very little loss of head. The mill jig installation is usually

FIG. 38. Denver Equipment Company unit flotation cell and hydraulic cone in tubemill classifier circuit.

a compromise between gold recovery and available space limitations, Although it is depended upon primarily for the removal of only the relatively coarse values, a 60 to 70 per cent recovery of the total free gold values at the usual mill grind is not at all unusual. A good example of jig operation is to be found in the Preston East Dome practice, a complete description of which will be found in C.MJ., Vol. 62, page 535. In placer treatment, the replacement of coco-matting-bottomed sluices by modern mechanical or placer-type jigs has resulted in marked improvement in recovery and lowered operating costs. The use of a hydraulically pulsated cleaner jig circuit in conjunction with continuous amalgamation has eliminated dredge shutdown time for cleanup and practically

CONCENTRATION

133

eliminated cleanup labor. Pilot sluices on jig tailings show almost complete saving of recoverable values by the jigs. Of interest is the comment of one South American operator that a proportion of the gold recovered by the jigs on his dredge is finer than can be recovered by native hand panners, The high ratio of concentration (200-1000: 1) made possible by proper jig adjustment enables the jig operator to strip off his values into a theft-

FIG. 39. Dorrco Pan-American jigs in grinding circuit. Corporation, Grass Valley, California.

(Idaho Maryland Mines

proof, locked hutch or a screened and locked area in which a cleaneramalgamation circuit can be operated. Jig Capacity. The capacity of a jig is a function of the bed area. The ratio of length of bed to width is an important factor in determining the recovery into the concentrate. In placer work a length to width ratio of from 2-4: 1 is used, being obtained by end flow grouping of a series of square cross-section cells. In mill circuit work space limitations usually restrict the length-width ratio to 1: 1 or 2: 1.

134


Dilution of the jig feed has a bearing on effective capacity; an adequate retention time through the machine must be obtained, and flooding with high-feed dilutions reduces overall capacity. Mechanically pulsated types of jigs such as the Dorrco Pan-American, the Bendelari, and the Denver Mineral Jig have an average capacity of 2 to 2.5 tons per square fobt of bed area per hour, expressed in terms of new feed. In mill circuits with high circulating loads the actual duty on the jig may be 10 tons per square foot per hour or more; it is noted that the correspondingly lower recovery of gold is accepted in such circuits. Water requirements for mechanically pulsated jigs vary widely from as low as 2 to as high as 10 gallons per minute per square foot of jig bed area. These variations depend on the dilution of the feed, the particular design of the jig, the size distribution of the feed, the grade of concentrate desired, and the speed and the stroke of the jig. These last factors are interdependent with the volume of hutch water added; it is usual to establish TABLE

17.

JIG STROKES AND SPEEDS

3 to 20 mesh

20 to 200 mesh

. .

% to 1 120 to 160

Va to % 160 to 350

. .

~~ to 1 250 to 400

~& to ~~ 400 to 600

Feed size

Mechanical jigs: Stroke, diaphragm, in Speed, strokes per minute Hydraulic jigs: Stroke, valve, in Speed, strokes per minute

"

"

the stroke and speed and then to use the volume of hutch water as an operating variable. Mechanical Conditions. The average California practice for stroke and speed is shown in Table 17. The screen and shot sizes depend on the size of the concentrate desired. The size of the voids determines the effectiveness of sifting. A coarse bed (e.g., ~-in. shot) recovers a coarse concentrate and has a larger capacity but may not exclude sand efficiently. A fine bed (e.g., %2-in. shot) recovers small grains effectively and produces a cleaner concentrate. The choice involves a knowledge of the size distribution of the ore minerals. CORDUROY AS A GOLD SAVER

The use of blankets, canvas, coco matting, and corduroy-in fact, any material with nap-to save gold not held by copper plates, also to catch some heavy materials, has persisted from the time of Agricola to the present. From this practice has corne the present use of corduroy alone to entrap the gold. It has replaced amalgamation on the Rand, is satis-

135

CONCENTRATION

factorily used in Australia and in Canada, and is slowly being introduced into the United States. Blankets are preferred at Kolar, India. The information on corduroy following is from the article by the late M. W. von Bernewitz in E. and lYIJ., February, 1935. Corduroy is a thick and durable cotton stuff or cloth, corded or ribbed=a pile or nap on a base. Corduroy for wearing apparel differs from that

(a)

(b)

FIG. 40. Weave and nap of corduroy, natural size.

(a) Plan; (b) end view.

136


used in gold recovery in having narrower and closer cords or ribs. Also, in industrial corduroy, the nap has been cut as in the pile on certain highgrade floor rugs, thereby offering innumerable spaces for entrapping gold. The material is known as a "pulp-sifting corduroy cloth." It is woven from special material, and after being put through a serrating machine, it has a deep pile or nap with strong backing. Then the cloth is shrunk. It is made up into bales of about 80 yd. in two widths-28 and 36 in. Figures 40a and 40b show the weave and nap of corduroy. Compared with the use of copper amalgamation plates, which they have almost completely replaced, corduroys show advantages in regard to lower

FIG. 41. Corduroy nap, enlarged ten times.

initial cost, less skill required to operate and maintain, no danger from mercurial poisoning, less possibility of theft of amalgam, smaller loss of time, fewer effects of bad water and scouring by coarse pulp, and necessary general supervision. Corduroy is not expensive, it does not require skilled labor, theft is not so simple, and a remarkably large flow of pulp can be passed over it. Even if an ore is ground fine, it may pay to use corduroy in the flow sheet and catch the larger particles of gold which might prolong cyanidation. What is known as "rusty" gold is amalgamable with difficulty, if at all. The metal may be coated with iron, occasionally with manganese or other elements. Copper plates will not catch this sort of gold, but corduroy will.

CONCENTRATION

137

Method of Applying Corduroy. In laying corduroy, the .cords or ribs are placed across the table and flow of pulp, with the high side of the nap facing the stream. Each length overlaps the succeeding length a few inches. ' In general, the slope of the corduroy table must be determined by actl~al trial-the size and specific gravity of the solids and the liquid-solid ratio being the principal factors that influence the slope. An unclassified pulp with up to 20-mesh sand at 4 or 5 to 1 dilution will flow down 1% in. per ft.; more dilute and finer pulp needs less slope. Roughly, 3 tons of ore passing over tile tables should have 1 sq. ft. of .,corduroy. Cloths are rinsed as often as may be necessary to keep the riffles from packing with heavy minerals. They are kept in use until the pile or backing gives way. The final handling of worn-out corduroy is to burn it and treat the ash separately. It should not be fed to the mill circuit for fear that the carbon may precipitate the gold from the cyanide solutions. At the Dome mill the corduroy is dumped loosely into a tank in which cyanide solution is circulated for several days. This dissolves any fine gold enmeshed in the fabric. Then the cloths are discarded. CANADIAN PRACTICE

At Dome Mine. The following excerpt describing the Dome blanket practice is from the paper by P. D. P. Hamilton in Trans. 112, A.I.M.E. In the Dome mill, destroyed by fire in 1929, free gold was removed from the circuit by a combination of amalgam plates and blankets. In the new mill an attempt was made to use an all-cyanide process, but this was quickly dropped, and a blanket plant installed to remove the free gold from the circuit. The feed to the blanket tables is the unclassified product (about 60 per cent minus 200 mesh) of the primary and secondary grinding circuits. The pulp is about 1.5 parts water to 1 part ore. The primary load is 1730 tons, with an additional circulating load of approximately 1100 tons, making a total load over the blankets of 2830 tons daily. There are 28 blanket tables of a combined area of 756 sq. ft., each table being 4 ft. 6 in. wide by 6 ft. long. The tables are of wood, set so that the inclination of the blankets is 1%: in. per ft. The corduroy blankets are 28 in. wide and cut 5 ft. long to allow for shrinkage during use. The first strip is held in place by a flat iron bar, but the other two strips are held in place by lapping the upper one over the next "lower one. The "special heavy-backed undyed corduroy 28 in. wide" is specially manufactured for this purpose by James Johnston, 18 London Road, Manchester, England. The strips are placed on the table with the high sides of the cords toward the flow of the stream, thus making the necessary rimes to catch the gold. The top blankets are washed every hour, the others every 2 hours, three men changing, washing, and looking after distribution' on tables, etc. When the blankets are changed, they are folded and rolled up so as to keep the product on the inside. They

138


are replaced by another set of blankets but before the next change are washed and rolled ready to be reused. The washing is done in boxes, one for each shift, the blankets being merely unrolled and moved longitudinally up and down in water two or three times. This removes the greater part of the product. When blankets are discarded, they are more carefully cleaned and are then treated by cyanide in a small tank until no further gold comes into solution. By means of a large rubber-lined rotary distributor, the pulp is delivered to seven smaller distributors, each serving four tables, from each of which the flow can be deflected to a return launder by a rubber-lined swing gate. Each wash tub has a pipe leading out of the bottom, so that it can be loaded directly into an amalgam barrel set directly below the tub. The amount collected in each tub in an 8-hr. shift is approximately 1 ton, which is the capacity of the amalgam barrels. The barrels are 36 in. in diameter by 5 ft. long and are lined with white-iron liners, but barrels with cast-iron shells are preferable because they can be used without linings. Linings invariably collect and hold up amalgam. Each barrel is charged with 200 lb. of H~-in. balls and is driven at 19 r.p.m. These barrels are loaded each day about noon, lime is added, and the barrels are closed and started. The real function of the grinding is to brighten all the gold so that it will amalgamate readily. At 6:30 the following morning the barrels are stopped and the requisite amount of mercury is added. The barrels are then rotated again for H~ hr. and then dumped. The mercury and amalgam are collected, cleaned up, and pressed into cakes which contain approximately 50 per cent mercury, carrying about 10 dwt. gold per ounce. The amalgam is taken to the refinery, where it is held until sufficient accumulates for a retort. In addition to the three men on each shift, two men on the day shift handle the dumping of the barrels and the cleaning of the amalgam. During the year ending Dec. 31, 1947, the total recovery at the Dome mill was 96.14 per cent, the blanket plant giving 63.30 per cent and the cyanide plant 32.84 per cent. RAND PRACTICE

The revival of corduroy originated on the Rand, where copper plates had saved many million of ounces of gold. About three-quarters of the Rand mills use corduroy, and the gold recovered varies from 24 to 69 per cent of the gold in the heads. Although written many years ago, Wartenweiler's paper in Jour.C.lJl. and 1J1.S.S.A., February, 1923, is one of the few and best references on corduroy. It is replete with figures and flow sheets and is entitled "Recovery of Gold by Blanket Concentration in Substitution of Plate Amalgamation." For one group of large gold producers, plates had saved 47 to 73 per cent of the gold. Plate amalgamation, however [to quote Wartenweiler], had become encumbered with a number of disabilities. . .. With the increasing practice of fine grinding, the perfection of cyanide extraction and of precipitation, the importance of high recovery by amalgamation had receded according to the degree of efficiency of the section of the reduction plant devoted to recovery of gold by cyaniding....

139

CONCENTRATION

The Apex plant, treating ore from the Modderfontein East, was among the first to discard plate amalgamation. The copper plates, from which accumulated amalgam had been removed, were converted to corduroy tables. No holding-down device was used, but air bubbles tliat formed were "ironed out." Even distribution of the pulp for corduroy is as important as for plates. Wartenweiler found the following after a 4-hr. run of pulp over a series of five corduroys: TABLE

18.

Corduroy

1 2 3 4 5

GOLD RECOVERED ON CORDUROY

.,

Gold Panned and Weighed, Per Cent

76 to 80 11 to 18 3.1 to 4.2 2.3to3.7 0.8 to 2.1

Current Use of Corduroy. The corduroy tables are generally placed immediately after the ball-mill or tube-mill discharge and are frequently used on both primary and secondary circuits. Thus, they have to handle the whole circulating load through the mill, and the pulp will carry 30 to 40 per cent of plus 90-mesh material. The average corduroy table is 5 ft. wide and 10 or 12 ft. long and has a slope of about 2 in. per ft. Approximately 1 sq. ft. of corduroy area is required per ton of ore milled per day, and about 1 ton of concentrate is produced per 1000 tons of ore milled. The corduroy strips are washed every 3 to 4 hr., and the pyrite concentrate is cleaned once a day on a Wilfley-type table. The cleaned concentrate is amalgamated in barrels, the amalgam being cleaned on bateas and retorted. RECOVERY OF GOLD AND SILVER ASSOCIATED WITH SULPHIDES QRAVITY CONCENTRATION

Though the methods previously discussed usually catch the coarser particles of sulphides in the ore and thus indirectly recover some of the gold associated with these and other heavy minerals, they are not primarily designed for sulphide recovery. Where a high sulphide recovery is demanded, flotation methods are now in general use, but in the days before flotation was known, a large part of the world's gold was recovered by concentrating the gold-bearing sulphides on tables and smelting or regrinding and amalgamating the product,

140


Though the modern trend is away from the use of tables, because flotation is so much more efficient, a brief description follows: As a matter of historical interest the following account of early operations at the Timmins Ochali Mining Co. property in Colombia is quoted:" Overflow from the batteries runs over Antioquenan tables which are flat and smooth and made of a fibrous wood known as "yolombo." They are arranged in series much the same as blanket or amalgamation plate tables. The grain of the wood in the deck runs with the flow of pulp, and a sharp-pointed tool is used to scratch its surface in a crosshatched pattern. The fibrous wood stands up along the scratches and forms an ideal trap for free gold. Twice a day the tables are cleaned and the high-grade concentrate is amalgamated by hand in a wooden batea or gold pan. Dresslnq water bOClrd",

.." C

w

F

o s: u

'"

Ci

TOIilineis Dischorrqe Siole (Por\en+eol)

)IW! . %' CnlSS

.,W

. FlowoFpu!p'

I)!Linoleum'_;'l'

~- ~.,~

Wood f"looring

i

Section of R-Iffles on

Cross Section of Malin Riffles on Prirnarv Concerrh-crlion Zone

IntermeoliOite PlolteolU

Wood

-F!oorin~

FIG. 42. Details of Deister concentrator.

During the early days of operation referred to, nine wooden mills totaling 66 stamps, one ten-stamp California-type stamp mill, and three sand-leaching cyanide plants were in operation. Use of this equipment was discontinued in May of 1936.

Bumping Tables. The gravity concentrators used in gold and silver milling plants are usually of the bumping-table or the endless-belt (Vanner) type. The former consists of a riffled deck carried on a supporting mechanism that permits adjustment of slide slope and connected to a "head" mechanism that imparts a rapid reciprocating motion in a direction parallel to the riffles. A cross flow of water is provided by means of launder or other type distributors mounted along the upper side of the deck. The feed enters just ahead of the water supply, concentrate is taken off at the lower end of the table, and tailings overflow the lower side (see Figs. 42 2

E. and M.J., Vol. 143, No.5, p. 58, May, 1942.

141

CONCENTRATION

and 43). A. M. Gaudin" in "Principles of Mineral Dressing," .1939, has in Chap. XIII made a comprehensive mathematical analysis of the principles involved in flowing film concentration and tabling. On p. 294 Gaudin states: The flowing film concentrators that are known as shaking tables utilize other principles besides those discussed so far (stationary tables). Shaking tables are provided with a reciprocating motion at right angles to the flowing fluid film and directed horizontally. This reciprocating motion of the table deck has an asymmetrical acceleration, the net effect of which is to cause intermittent travel of solids resting on the table. This is one of the auxiliary principles utilized in shaking tables. oJ

FIG. 43. Stratification of mineral on a Wilfley table. Another auxiliary principle derives from the use of riffles which disturb the viscous flow of the fluid across the deck and substitute for it a fluid composed of one top layer flowing more or less by viscous flow and an eddying bottom layer. Riffles are of great importance to tabling. They are responsible for the increased capacity of riffled decks over smooth ones. This is realized when it is considered that a smooth deck treats a bed one particle deep and a riffled deck treats a teetering suspension of particles often many particles deep....Each trough between successive riffles is a place where hindered settling and consolidated trickling occur.

The capacity of tables varies widely between a maximum of, perhaps. 200 tons per 24 hr. in coarse roughing operations to 5 to 10 tons per 24 hr, on very fine feeds. Around 10 mesh is about the upper practical size of 3

Professor of Mineral Dressing, M.LT., Cambridge, Mass.

142


feed, while only incomplete recoveries are made in the 200 to 325 mesh range. Finer material is lost on all but very low-capacity slime tables. Vanners. The endless-belt or Vanner type of table was developed for the purpose of recovering the fine mineral particles lost in ordinary table operation. With a feed range of minus 65 to 100 mesh, particles as fine as 10 to 20 microns are recovered, but the capacity of such tables is only 1 to 3 tons per 24 hr., and the vanner is now practically obsolete. Tables can be operated on either classified (sized) or unclassified feed. The use of an unsized feed, particularly in roughing operations, as at Hollinger, is common practice, but it is generally recognized that preliminary sizing of the feed with each sized band going to a separate table is the more efficient practice. A. W. Fahrenwald discusses this point in T.P. 403, 'U.S.B. of M., 1927. He concludes that preliminary classification, particularly if classifiers of the hydraulic types are used, gives increased recovery, a higher grade concentrate, greater table capacity, and less middling for regrind; in other words, classification enables a table to do its best work. The Dorrco sizer described in Chap. VI, the Richards pulsator classifier, and others employing classifying zones of uniform cross section give efficient results. The Humphreys Spiral Concentrator. This device, which was invented by 1. B. Humphreys and first used in 1943 for concentrating chromite in Oregon beach sands, consists of five or six spiral turns of a modified semicircular launder which is about the size of a conventional automobile tire. Feed enters the top spiral and the tailing discharges from the bottom one, while concentrate and middlings are cut off by outlet ports regularly spaced at each turn of the spiral, and the products passed through rubber hoses to common launders which run the full length of a bank of spirals. Wash water is supplied from a small wash-water channel paralleling the main channel. Operating entirely by gravity flow and involving no mechanical parts, the separation of the heavy constituents of the feed is effected by the same centrifugal forces and flow gradients encountered in ordinary river or stream concentration. A capacity of 38 tons per spiral was obtained in the 1000-ton per 24 hr. Oregon plant operating on about a minus 40-mesh feed and in the 5000ton plant recently installed near Jacksonville to concentrate ilmenite 174 roughing, and 12 finishing spirals have replaced an installation of tables and flotation cells. Although we do not know of Humphreys spirals being used to concentrate gold ore, they should be effective for the recovery of free gold such as in the treatment of placer sands.

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143

FLOTATION

The flotation process, which is today so extensively used for the concentration of base-metal sulphide ores and is finding increased use in many other fields, was introduced commercially into the United States in 1911. While its development was rapid and by 1932 some 60 million tons a year' was being treated in 268 plants in the United States alone, it was not until about this same year that flotation plants began to be installed for the treatment of gold and silver ores as a substitute for or in conjunction with cyanidation.

FIG. 44. Denver Sub-A flotation machines, showing the different multiple-cell units available to meet a wide range of tonnage requirements.

The principles involved and the rather elaborate physicochemical theories advanced to account for the selective separations obtained are beyond the scope of this book. Suffice it to say that in general the sulphides are air-filmed and "floated" to be removed as a froth from the surface of the pulp while the nonsulphide "gangue" remains in suspension, or "sinks," as the expression is, for discharge from the side or end of the machine. For more complete information reference is made to Taggart's H andbook of Mineral Dressing, 1945; Gaudin's Flotation and Principles of Mineral Dressing; 1. W. Wark's Principles of Flotation; and the numerous papers on the subject published by the A.I.M.E. and U.S. Bureau of Mines. FLOTATION MACHINES

Flotation machines can be classed roughly into mechanical and pneumatic types. The first employ mechanically operated impellers or rotors

144


for agitating and aerating the pulps, with or without a supplementary compressed-air supply. Best known of these are the Mineral Separation, the Denver, the Fagergren, the Agitair, and the Massco-Fahrenwald. Pneumatic cells use no mechanical agitation (except the MacIntosh, now obsolete) and depend on compressed air to supply the bubble structure and to hold the pulp in suspension. Well-known makes include the TABLE

Number

Size, in.

19.

FLOTATION MACHINE DATA*

Cell volume, cu. ft.

Actual hp,

per cell

Motor size for two cells

Approx. capacity, t dry tons per 24 hr.. sp. gr. 2.8 at 25% solids,

4-cell

I fi-cell I S-cell

\ lO-ee11

Denver flotation machines 8 12 15 18 18 S.P. 21 24 30

16 by 22 by 24 by 28 by 32 by 38 by 43 by by

'56

16 22 24 28 32 38 43 56

0.75 1.0 1.2 1.4 2.2 3.2 4.2 9.0

3.0 10.0 12.0 18.0 24.0 40.0 50.0 100.0

1.5 2.0 3.0 3.0 5.0 7.5 10.0 20.0

16 45 65 95 120 215 250 535

24 70 95 i40 180 320 380 800

32 95 125 185 240 425 500 1070

40 115 160 235 300 535 630 1335

Fagergren flotation machinesj

Size, in.

Cell volume, cu. ft.

36 44 56 66

12 21 42 62

Motor hp,

3 5 10 15

Capacity per row of machines, tons per 24 hr.

50 to 250 to 500 to 1000 to

250 500 1000 2000

* From

manufacturers' catalogues. For medium floating ores = about 10 min. flotation times. +The machines are built in one-, two-, three-, or four-cell units, and in the case of multicellular construction, the individual cells are separated from each other by a partition plate, but the ore pulp passes directly through suitable openings from one cell to the next.

t

Callow and MacIntosh (no longer manufactured) the Southwestern, and the Steffensen, the last, as shown in the cross-sectional view in Fig. 47, utilizing the air-lift principle, with the shearing of large bubbles as the air is forced from a central perforated bell through a series of diffuser plates. Calculating Flotation-cell Requirements. The number and size of flotation cells required for any given installation are readily determined

145

CONCENTRATION

if the problem is looked upon as a matter of detention time for a certain total volume of pulp. The pulp flow in cubic feet pel' minute is determined from the formula

C =

~ + W) ( sp. gr.

X 32

1440 where C = T. = sp.gr. = TV =

cubic feet per minute tons dry solids per 24 hr. in feed specific gravity of solids tons of water per 2"1 hr. in feed

oJ

"'"-FIG. 45. Mineral-bearing froth being discharged into the concentrate launder of a four-cell No. 18 Denver Sub-A flotation machine. (Denver Eqltipm~nt Company.)

Having obtained the volume to be handled per minute," the total cell volume to be provided is found by multiplying this volume by the contact time in minutes required in the flotation machines. For ordinary ratios of concentration the effect oncell capacity of concentrate (or froth) removal can be neglected, but where a high proportion of the feed is taken off as concentrates, or where middlings are removed for retreatment in a separate circuit, due allowance should be made for reduced flow and, in consequence, increased detention time toward the tail end of a string of cells. Not less than a series of four cells and preferably six or more cells should be used in any roughing section in order to prevent short-circuiting.

146

CYANIDATION AND CONCENTRATION OF ORES FLOTATION REAGENTS

It is not intended here to discuss the subject of flotation reagents in any detail. The subject is a large one with a comprehensive technical and patent literature. Research leading to the development of new reagents and to our understanding of the mechanism involved has been largely in the hands of academic institutions and the manufacturers of chemical products. Recent work reported by A. M. Gaudin! on the use of "Radioactive Tracers in Milling Research" described, for instance, the use of a flotation reagents containing radioactive carbon to determine the extent of collector adsorption. The "bubble machine" devised to measure the angle of contact of air bubbles on collector-treated mineral surfaces has been extensively used for determining the theoretical value of various reagents as flotation collectors, but for the most part the actual reagent combination in use in commercial plants is usually the result of trial-and-error methods. The following is a brief discussion of the reagents ordinarily used for the flotation of gold and silver ores prepared from notes submitted by S. J. Swains on and N. Hedley of the American Cyanamid Company. REAGENTS USED IN FLOTATION OF GOLD ORES

Conditioning agents are commonly used, especially when the ores are partly oxidized. Soda ash is the most widely used regulator of alkalinity. Lime should not be used because it is a depressor of free gold and inhibits pyrite flotation. Sodium sulphide is often helpful in the flotation of partly oxidized sulphides but must be used with caution.because of its depressing action on free gold. Copper sulphate is frequentlyrhelpful in accelerating the flotation of pyrite and arsenopyrite. In rare instances sulphuric acid may be necessary, but the use of it is limited to ores containing no lime. "Ammo-phos," a crude monoammonium phosphate, is sometimes used in the flotation of oxidized gold ores. It has the effect of flocculating iron oxide slime, thus improving the grade of concentrate. Sodium silicate, a dispersing agent, is also useful for overcoming gangueslime interference. Promoters or Collectors. The commonly used promoters or collectors are "Aerofloat" reagents and the xanthates. The most effective promoter of free gold is Aerofloat flotation reagent 208. When auriferous pyrite is present, this reagent and reagent 301 constitute the most effective promoter combination. The latter is a higher xanthate which is a strong and nonselective promoter of all sulphides. Amyl and butyl xanthates are also widely used. Ethyl xanthate is not so commonly used as the higher xanthates for this type of flotation. 4

Annual meeting, C.I.M. and M., Montreal, Quebec, March, 1949.

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147

The liquid flotation reagents such as Aerofloat 15, 25,and 31 are commonly used in conjunction with the xanthates. These reagents possess both promoter and frother properties. When malachite and azurite are present, reagent 425 is often a useful promoter. This reagent was developed especially for the flotation of oxidized copper ores. The amount of these promoters varies considerably. If the ore is partly oxidized, it may be necessary to use as much as 0.30 to 0.40 lb. of promoter per ton of ore. In the case of clean ores, as little as 0.05 lb. may be enough. The promoter requirement of an average ore will usually approximate 0.20 lb.

FIG. 46. Sectional view of a Fagergren flotation cell.

Frothers and Froth Modifiers. The commonly used "frothers" are steam-distilled pine oil, cresylic acid, and higher alcohols. The third mentioned, known as duPont frothers, have recently come into use. They produce a somewhat more tender and evanescent froth than pine oil or cresylic acid; consequently they have less tendency to float gangue, particularly in circuits alkaline with lime. The duPont frothers are highly active frothing agents; therefore it is rarely necessary to use more than a few hundredths of a pound per ton of ore. When coarse sulphides and moderately coarse gold (65 mesh) must be floated, "froth modifiers" such as Barrett Nos. 4 and 634, of hardwood

148


creosote, are usually necessary. The function of these so-called froth modifiers is to give more stable froth having greater carrying power. REAGENTS FOR SILVER ORES

The conditioning agents used far silver ores are the same as those for gold ares. Soda ash is a commonly used pH regulator. It aids the flotation of galena and silver sulphides. When the silver and lead minerals are in the oxidized state, sodium sulphide is helpful, but it should not be added until after the sulphide minerals have been floated, because sodium sulphide inhibits flotation of the silver sulphide minerals. Aerofloat 25 and 31 are effective promoters for silver sulphides, sulphantimonites, and sulpharsenites, as well as for native silver. When galena is present, No. 31 is preferable to No. 25 because it is a mare powerful galena promoter. Higher xanthates, such as American Cyanamid reagent 301 and amyl and butyl xanthates, are beneficial when pyrite must be recovered. When the are contains oxidized lead minerals, such as anglesite and cerussite, sodium sulphide and one of the higher xanthates may be used. In some instances reagent 404 effects high recovery of these minerals without the use of a sulphidizing agent. Silver ores require the same frothers as gold ares-viz., pine oil, cresylic acid or duPont frothers. "Aero," "Amrno-phos," and "Aerofloat'" are registered trade-marks applied to products manufactured by this company. The Great 'Western Electro-Chemical Company, California, makes amyl xanthate, butyl xanthate, potassium xanthate, and sodium xanthate. In the United States these reagents are used on the gold ores of California and Colorado and in Canada on the gold ores and sulphides of Ontario and Quebec. Flotation reagents of the Naval Stores Division of the Hercules Powder Company are as follows: Yarmor F pine oil, a frother for floating simple and complex ores; Risor pine oil, for recovering sulphides by bulk flotation; Tarol 1, a toughener of froth, generally used in small amount with Yarmor F, but with some semioxidized ores where high recovery is essential yet the grade of concentrate not so important, Tarol does good work; Tarol 2, a frother for floating certain oxide minerals, but it can be used in selective flotation of sulphide minerals and in bulk flotation where tough froth is desirable; Solvenol, for the floating of graphite in conjunction with Yarmor F. PRECIPITATING EFFECT OF FLOTATION REAGENTS

The statement has come to the attention of the American Cyanamid Company that organic flotation reagents, such as xanthates, even in the

CONCENTRATION

149

small amounts used in flotation, cause reprecipitation of gold from pregnant cyanide solutions. The ore-dressing laboratory of this company is studying the question, and preliminary results indicate that this statement is unfounded. The addition of xanthate, in the amount usually found in flotation circuits, does not precipitate gold from a pregnant cyanide solution containing the normal amount of cyanide and lime. PRIMARY SLIME

Valueless slime, in addition to its detrimental effect in coating goldbearing sulphide, thereby limiting or preventing its flotation, also becomes mixed with the flotation concentrate and Iowers its value. Sometimes the problem in flotation is that, although the gold is floatable, the concentrate product is of too low grade. Talc, slate, clay, oxides of iron, and manganese or carbonaceous matter in ores early form slime in a mill, without fine crushing. Such "primary slime," according to E. S. Leaver and J. A. Woolf of the U.S. Bureau of Mines, interferes with the proper selectivity of the associated minerals and causes "slime interference." The tendency of primary slime is to float readily or to remain in suspension and be carried over into the concentrate. Preliminary removal and washing of this primary slime before fine crushing is one method of dealing with it. At the Idaho-Maryland mill, Grass Valley, Calif., starch is regularly used as a depressant during flotation. Flotation tests using starch were made on a quartz ore containing carbonaceous schist from the Argonaut mine, Jackson, Calif.; a talcose ore from the Idaho-Maryland mine mentioned; a talcose-clayey ore from Gold Range, Nev.; a siliceous, iron and manganese oxide ore from the Baboquivari district, Nevada; carbonaceous and aluminous slime from the Mother Lode and some synthetic ores. The conclusions from the foregoing tests were in part as follows: 1. Finely divided metallic gold in milling ores floats readily, and a high-grade concentrate can be made by flotation if no interfering slime or gangue is present. Any good collector may be used for the flotation of gold, but organic collectors of the xanthate type produce a cleaner, higher grade concentrate than coal tarcresote oils. 2. Some "protective colloid" should be added to "wet out" talcose or carbonaceous slime and destroy its tendency to float. 3. Clayey slime does not have strong, fictative properties, but it tends to remain in suspension and coat mineral particles, making it difficult to obtain good selectivity during flotation. Proper deflocculation of an ore pulp and some agent to destroy the flotutive property of clay improve flotation of this type of ore. 4. It is essential to keep pulp containing iron (not hematite) and manganese oxides in a dispersed condition, and improved results are obtained by means of a depressing agent such as starch. 5. Starch was the most effective depressing agent tried. It should be added as a solution to the ore pulp. Starch displays a selective action in its depressing effect.

150


It acts first on the slime; then, if a sufficient excess of starch is present, it will cause some depression of sulphides and metallic gold, either by wetting out or by producing an extremely brittle froth. Therefore, care must be taken in regulating the amount of starch added to obtain the maximum depression of the slime commensurate with high recovery of the gold. In this, as in all other phases of flotation, each ore presents an individual problem and must be so studied.

The American Cyanamid Company in "Ore Dressing Notes," BuZ. Number 9, January, 1939, describe the use of their 600 series of reagents which were developed primarily for the purpose of depressing carbonaceous and siliceous slimes in the flotation of-gold ores. Carbonaceous material not only greatly increases the bulk and moisture content of a flotation concentrate, but its presence makes cyanidation of the concentrate difficult or impossible owing to reprecipitation of the gold during treatment. In the treatment of an auriferous sulphide ore associated with carbonaceous shale from South Africa, up to 77 per cent of the carbon was eliminated by the use of 1 lb. per ton of reagent 637 with a' 90.5 per cent gold recovery at 20.4: 1 ratio of concentration. A gold carbonaceous sulphide ore from California carrying free gold yielded a 93 per cent recovery into a concentrate at 14.4:1 to ratio of concentration after conditioning with 0.50 lb. per ton of reagent 645. In each case the ore was ground to about 70 per cent minus 200 mesh and conditioned at 22 per cent solids with the reagents as indicated. Flotation reagents included reagents 301 and 208 and pine oil. In the second case some soda ash and copper sulphate were also used. FLOTATION IN THE FLOW SHEET

It is obvious that the most suitable treatment for ores carrying gold and silver associated with pyrite and other iron sulphides, arsenopyrite or stibnite, will depend on the type of association. Cyanidation is usually the most suitable process, but it often necessitates grinding ore to a fine size to release the gold and silver. Where it is possible to obtain a good recovery by flotation in a concentrate carrying most of the pyrite or other sulphides, it is often more economical to adopt this method, regrinding only the comparatively small bulk of concentrate prior to the leaching operation. That the trend over the last 10 years has been in this direction will be noted from the numerous examples of such flow sheets in Canada and Australia (see Chap. XV). A number of plants formerly using all-cyanidation have converted to the combined process. The suitability of the method involving fine grinding and flotation with treatment of the concentrate and rejection of the remainder should receive careful study in the laboratory and in a pilot plant. McIntyre-Porcupine

CONCENTRATION

151

ran a l50-ton plant for a year before decidin~ to build its 2400-ton mill. Comparative figures given by J. J. Denny in E. and lJ!LJ., November, 1933, on the results obtained by the all-sliming, C.C.D. process formerly used and the later combination of flotation and concentrate treatment showed a saving of 12.1 cents pel' ton in treatment cost and a decrease of 15 cents per ton in the residue, a total of 27.1 cents per ton in favor of the new treatment. Flotation may also prove to be the more economical process for the ore containing such minerals as stibnite, copper-bearing sulphides, tellurides,

FIG. 47. Sectional view of the Steffensen pneumatic flotation machine.

and others which require roasting before cyanidation, because this reduces the tonnage passing through the furnace. T. B. Stevens in },I!.111., October, 1933, discusses the treatment at Lake View and Star in Australia. The total cost of all-sliming and flotation treatment of 45,000 tons a month in 1933 was 6.3s ($1.52) per ton, compared with 16s ($3.84) when dry crushing all the ore, roasting, and cyaniding 15,000 tons a month. Even when recovery of gold and silver from such ores by flotation is low, it may be advantageous still to float off the minerals that interfere with cyanidation, roasting, and leaching or possibly to smelt the con-

152


centrate for extraction of its precious metals. Cyanidation of the flotation tailing follows, this being simpler and cheaper because of prior removal of the cyanicides. Flotation Unit in the Grinding Circuit. It is a good practice to recover as much of the gold and silver as possible in the grinding circuit by amalgamation, corduroy strakes, or other gravity means to prevent their accumulation in the classifier; otherwise gold that is too coarse to float may escape from the grinding section into the flotation circuit where it will pass into the tailing and be lost. To prevent this, several companies including the McIntyre-Porcupine at Timmins, Ontario, have inserted a combination of flotation cell and hydraulic cone in their tube-mill classifier circuits. At the MclntyrePorcupine, according to J. J. Denny in E. and M.J., November, 1933, this cell is a Denver 500 Sub-A type. The total pulp discharged from each tube mill passes through 4-mesh screens which are attached to the end of the mills. The undersize goes to the flotation cell, and the oversize to the classifiers. Tailing from the cell flows to the classifiers, and the flotation concentrate joins the concentrate stream from .the main flotation circuit. The purpose of the hydraulic attachment is to remove gold that is too coarse to float, thus avoiding an accumulation in the tube-mill circuit. The cones have increased recovery from 60 to 75 per cent. Every 24 hr. the tube-mill discharge is diverted to the classifiers. Water is added for 15 min. to separate the gangue in the cells from the high-grade concentrate, after which a product consisting of sulphides and coarse gold is removed through a 4-in. plug valve equipped with a locking device. Each day approximately 400 lb. of material worth $2000 to $3000 is recovered. This is transferred to a tube mill in the cyanide circuit, with no evident increase in the value of the cyanide residue. The object of this arrangement is, of course, primarily to deplete the circulating load of an accumulation of free gold and heavy sulphides. Flotation of Cyanide Residues. Flotation is used to recover residual gold-bearing sulphides and tellurides. The Lake Shore mill retreatment plant is an interesting example of this technique (see Chap. X, page 159). The problem here was, of course, to overcome by chemical treatment the depressing action of the alkaline cyanide circuit on the sulphides. A full discussion of this and of the somewhat controversial subject as to whether flotation should in such an instance be carried out before, or after cyanidation will be found in J. E. Williamson's paper "Roasting and Flotation Practice in the Lake Shore Mines Sulphide Treatment Plant" elsewhere referred to. Summing up the specific considerations governing the choice of treatment, the author says:

CONCENTRATION

153

1. Virtually all exposed gold and telluride values in the ore min be safely and cheaply extracted by cyanidation of the raw ore.' . 2. The roasting of an uncyanided flotation concentrate made from the raw ore did not produce a product suitable for cyanidation. 3. Cyanidation of a flotation concentrate made from the raw ore, though it would clean up the exposed values, was a difficult and costly treatinent. 4. The development of a cheap and practical method of conditioning the cyanide tails made it feasible to concentrate the values contained in the sulphides and treat them at a profit by roasting. Incidental matters that influenced the choice of treatment scheme included the realization that preliminary flotation would have involved two separate treatment circuits with additional steps of thickening and filtration following the flotation. Furthermore, in the conditioning method evolved, as much as 60 per cent of the dissolved values in the cyanide tailings were precipitated and recovered.

There are, however, cases where flotation equipment was put in for the purpose of recovering the gold in a concentrate and rejecting the tailing only to find that the tailing was too valuable to waste and had finally to be cyanided before discarding. It is generally true that cyanidation is capable of producing a tailing of lower gold content than flotation. At a price of $35 pel' ounce for gold this fact is of much greater importance than when gold was valued at $20.67 per ounce. The possible gold loss in the residue to be discarded will influence the choice of a method of treatment. CONCENTRATE TREATMENT

In the case of free-gold recovery, the small bulk of gravity concentrate obtained is usually treated by barrel amalgamation on the property (see Chap. XI). In the treatment and disposal of flotation concentrates both metallurgical and economic factors must be considered. In general, local treatment is preferable, but proximity to a smelting plant, low transportation rates, and reasonable smelting charges may favor the shipping of concentrates over making the capital outlay for a local plant and treatment. For local treatment of gold-bearing sulphides these metallurgical questions must also be answered: Shall the concentrate be treated raw or The following interesting observation is made in this connection: It has been pointed out by millmen in other camps treating ores containing both sulphides and tellurides that, as long as cyanidation is carried out on the raw ore, where the concentration of sulphide and telluride is low, good recovery of value in tellurides is possible. However, once the sulphides and tellurides are concentrated, it is next to impossible to extract the gold tellurides fully by any commercially feasible cyanide treatment. 5

154


roasted? Shall the concentrate residue be discarded or returned to the main pulp flow for further treatment? The prevailing practice in any district will depend upon such factors as the mineral association and local cost of treatment. In Australia the usual procedure is to treat concentrates by roasting and cyaniding; on the Rand treatment of the raw concentrate is preferred; in North America both schemes are used depending upon the particular ore being treated. Raw treatment of concentrates usually requires fine grinding and prolonged contact-up to 16 days-with cyanide solutions. Concentrates in which the gold is associated only with pyrite usually respond satisfactorily to raw treatment by cyaniding following fine grinding; when the gold is associated with other minerals, raw treatment frequently does not give satisfactory extraction. Treatment after roasting involves attrition milling or agitation to break up lumps (sometimes water washing) and finally cyanidation (see "Cyanidation of Concentrates," American Cyanamid Company, Ore Dressing Notes 6; also Chap. X, "Roasting"; see also Fig. 75 for an example of typical concentrate treatment). NORTHERN ONTARIO PRACTICE

Practice in "handling" sulphides in the Kirkland Lake and Porcupine districts varies from grinding and treating them concurrently with the remainder of the ore to entirely separate treatment and is summarized as follows for the plants visited: Kirkland Lake District. Teck-Hughes and Toburn grind the ore as a whole and cyanide it by agitation and filtration and by decantation and filtration, respectively. Lake Shore floats its sulphides and tellurides after the whole pulp has been cyanided, then regrinds the minerals, roasts and cyanides them again. Separate cyanide circuits are maintained for the regular mill and the concentrate-treatment plant. Wright-Hargreaves floats its sulphides after the whole pulp has been cyanided, regrinds the minerals, and returns them to the circulating pulp for further treatment. There is no accumulation of sulphides in the circuit. Porcupine District. Coniaurum grinds the ore as a whole and cyanides it by agitation and filtration. Dome, by means of cone classifiers, builds up a circulating load of sulphides in the pulp from 3.75 to 9 per cent. These are continuously ground, passed over corduroy tables, and mixed with the remainder of the pulp for agitation and filter pressing.

CONCENTRATION

155

Hollinger collects its sulphides on Deister double-deck tables, regrinds and cyanides them, and then mixes them with the remainder of the pulp for further cyanidation and filtration. ]1.1clntyre-Porcupine floats its sulphides and cyanides them only. The remainder of the pulp is discharged to waste. SMELTER TREATMENT

In regard to shipping ores and concentrates to custom mills or smelters for treatment, Taggart? discusses the subject of treatment charges and returns in some detail. The following notes from the revised article on "Selling Metallic Ores and Concentrates" by P."M. Tyler will be of interest: Custom mills are built to treat a composite of the ores in the district which they serve. The district is usually small in area, because freight quickly eats up all possible profit on shipments of raw ore, despite the fact that tariffs are normally lower for low-grade ores. Recoveries tend to be lower in custom mills than in company mills, owing to changing character of feed, even when elaborate facilities for mixing feeds are available; this difficulty is, of course, accentuated when ores are run through individually. Many smelters maintain custom mills. Milling charges depend upon the kind of ore, the size and frequency of shipment, the extent of segregation required, whether the ore is purchased by the milling plant, whether the mill is at a smelter, the number of unusual conditions in the transaction, and the extent of competition. Most custom mills are for treatment of gold- and silver-bearing ores. Smelters have a more difficult problem in the purchase of base-metal ores with or without a procious-metal content; consequently, their schedules of charges and methods of payment are more complicated than those of custom mills. The smelter buys ores on the basis of the agreed assay, paying for valuable metals contained therein at prices current in principal metal-market centers, either at date of purchase or at some agreed date thereafter meant to be the probable date of sale, less a charge covering the cost of treatment and profit thereon. The treatment charge must include the cost of delivering ore to the smelter, sampling, smelting, freight on crude metal to the refinery, refining, selling, and a carrying charge on metal from the time of purchase to the time of disposal. Various methods of assessing these charges and the profit on operations are followed. In the case of some metals all is included in a treatment charge; in other cases a part only, viz., smelting, is included in the base treatment charge, the balance being taken care of in the price at which metal is paid for after certain deductions from the market price. All methods have as the fundamental basis of charge the cost of the items above enumerated. Smelter open schedules are published tenders by the smelter to purchase or treat ores and concentrates under stated conditions as to price and other items. The elements are (1) the treatment charge; (2) penalties; (3) payments. Treatment charge covers the actual cost of treating the ore plus interest on capital and investment, etc., and allows for working profit. Charges tend to graduate upward for low-grade material. 1

Taggart, op. eit., Sec. 2-255 et seq.

156


Penalties are imposed for ore constituents, which add to the difficulties and costs of smelting. They are normally graduated according to the content of unwanted material or the deficiency of desired material, Penalties and bonuses are quoted us so much a unit, which is 1 per cent or 20 lb. per short ton (22.4 lb. per long ton). Crude gold or silver ores shipped to smelters are typically siliceous; concentmtes, on the other hand, generally contain an excess of iron. Ores Lhat cont.ain too little Ieud or copper to serve us collectors of the precious metals in smelting are termed "dry ores." Highly siliceous ores (60 per cent upward of SiO o) are valued by copper smelters for use in converters, whereas they are highly penalized by lead smelters. In certain districts, however, siliceous ores are smelted on a flat schedule designed to encourage mining and maintain a flow of ore to the smelter. In Leadville, Colo., the treatment charges per ton are scaled from $5 for such ores worth $14 or less a ton to $10 for ores worth $50 or more per ton. Payments for a given metal differ materially according to the kind of smelter buying, und, of course, to the character of material shipped. Thus a lead smelter may pay about 4 cents per pound less for copper in a given ore or concentrate than a copper smelter, but the latter will pay for only half the lead at 2 cents less per pound. Payments by the smelter are rarely, if ever, bused on the full amount of a given metal in the product shipped, as shown by assay, or on the full market value of the metal at the time of settlement. The first difference is supposed to take care of losses in treatment. The second deduction is to cover freight on base bullion and the cost of refining it, the cost of recovery of by-products, selling, and it serves also as a hedge on the course of the market between settlement date and sale. Payment deductions and penalties are, in most cases, the sources of smelter profits. Gold payments, formerly based on $20.67 per ounce, have been based since 1933 on the realized mint price of $34.9125 per ounce ($35 less $0.0875 refining charge). Both lead and copper smelters usually pay for all Au over about 0.03 ounce per ton; the minimum ranges from 0.02 to 0.05; some contrncts deduct the minimum from the settlement assay. The price commonly varies according to the Au assay. Silver payments have generally been based in North America on the New York (Handy and Harman) quotations for the week during which the last car of the lot arrives at the smelter, but under the Silver Purchase Act domestically mined silver (accompanied by the necessary affidavits) is paid for on the basis of the realized mint price. Smelters usually deduct 1 oz. (sometimes only 0.5 oz.) per ton from the assay and pay for 95 per cent of the balance, sometimes making an additional deduction on nondomestic ores from the market price. Zinc and antimony smelters often make no payment for Ag or Au, and those which do deduct heavily because their losses in recovering precious metals are higher than those of copper or lead smelters. Native placer and bullion from amalgamation or cyanidation plants can be deposited at the United States Assay Offices in New York and Seattle or at the mints at Philadelphia, San Francisco, Denver, and New Orleans and will be paid for usually within 3 to 5 days at the market price of these metals less specified>refining charges, provided the gold or gold and silver content is 20 per cent or more. Silver, free from gold, will not, necessarily, be accepted unless needed for coinage or under some special provisions (e.g., the Silver Purchase Act). No allowance is made for platinum or base metals contained in bullion.

Ernest Gayford, in Trans. 112, A.I.M.E., 1934, in discussing ore treatment as a factor in small gold-mining enterprises (15 to 150 tons per day),

CONCENTRATION

157

at the higher price ($35 pel' ounce) for gold, concludes that, as with base metals, the nearer the producer can get to making a finished product himself the greater will be his reward, the net returns for bullion being necessarily higher than for the same amount of gold sent to a' smelter in concentrates. On a shipment of bullion containing 100 oz. fine gold, the return would be $34.81 an ounce. If 50 tons of 2-oz. concentrate, 20 tons of 5-oz. concentrate, or 10 tons of lO-oz. concentrate were sent to a smelter, the net respective yields based on average conditions would be $26.03, $28.94, and $30.64 pel' ounce. Of course the charges against the concentrates include haulage, freight, and treatment"

CHAPTER X

Roasting In the case of many complex gold and silver ores roasting before cyanidation is essential if satisfact01'y extraction of the precious metals is to be obtained. In such cases no practical amount of grinding or prolonged contacting of the raw ore with cyanide eoiutior. will effect more than a certain low extraction; in other cases, while the extractions may be acceptable, the consumption of cyanide is prohibitive, and 1'Oasting or, occasionally, pyritic smelting of the material is the only alternative. Roasting, which is today almost always carried out following concentration of the ualue« into a small bulle, must be used if it becomes necessary to decompose the minerals with which the gold is associated in order to expose it to solvent action. 111any of the cyanicides in the raw ore are broken up and rendered harmless by roasting, but the calcine may contain new com.poumde that must be removed by water 01' acid washing before cyanide treatment can be successf1llly carried out. GENERAL DISCUSSION

There were instances, particularly in Australia, where at one time the whole of the ore was roasted before cyanidation, but this practice has been for the most part discontinued on account of the large plant and high over-all costs involved. With improvements in flotation methods, the generally accepted procedure today is to concentrate the gold- or silverbearing minerals into a small bulk and roast the concentrate only. The calcine is then usually treated in a small cyanide circuit, with the residues sometimes passing into the main cyanide circuit, if the flow sheet includes such a step, or sent directly to waste. Those types of gold ores which most frequently require roasting include ores carrying arsenopyrite, stibnite, sulphotellurides, and pyrrhotite. Straight pyritic ores, where the pyrite is present in small quantities, usually yield their gold to fine grinding and cyanidation alone. Silver ores containing the values as polybasite, stephanite, pyrargyrite (the antimony sulphide), and proustite (arsenic sulphide) usually require roasting. Tetrahedrite is often refractory even after roasting. Argentite (the silver sulphide) and cerargyrite (the chloride) can frequently be cyanided without roasting. A comprehensive description of roasting practice is to be found in a series of articles entitled "Roasting Gold-Silver Sulphide Ores and Concentrates" by lVI, W. von Bernewitz appearing in C.1I1.J. in 1940. 158

ROASTING

159

A more recent paper, "Roasting of Arsenical Gold Ores and Concentrates" by F. R. Archibald presented at the Annual meeting of the C.I.M. and M. in Montreal, April, 1949, contains a bibliography listing 23 references on the subject of roasting gold and silver ores. There are four general types of roasting furnace in use: (1) the horizontal, multiple-spindle type, of which the Edwards furnace and modifications of it are typical; (2) the rotary kiln; (3) the vertical, multiple-hearth furnaces with rabble arms attached to a central shaft, of which the Wedge furnaces are typical; and (4) the Dorrco FluoSolids reactor. PYRITE

The roasting of straight pyritic ores involves the conversion of the iron sulphides to the oxide under oxidizing conditions with the evolution of sulphur dioxide and, to some extent by catalytic action, sulphur trioxide gas also. In this chemical change the iron mineral is rendered more or less porous, thereby permitting the dissolution of the contained gold by subsequent cyanidation. Though simple in principle, considerable temperature and other controls are necessary to avoid undesirable side reactions which are discussed below. It is not usual, for instance, to leave more than about 0.1 to 0.15 per cent insoluble sulphur in the roasted ore, but much larger percentages of soluble sulphates are often present. It is believed that in certain cases such salts may reduce the porosity of the iron oxide and so lower the gold extraction. In addition to higher gold losses, incomplete roasting also causes trouble due to the presence of ferrous salts and other cyanicides. Lake Shore Mines, Ltd. The treatment scheme at Lake Shore includes direct cyanidation followed by flotation of the sulphides, roasting of the concentrates, and cyanidation of the calcine in a separate circuit. The following notes are taken from J. E. Williamson's paper "Roasting and Flotation Practice in the Lake Shore Mines Sulphide Treatment Plant," C.I.M. and M. The pyrite values include all gold so intimately associated with the pyrite that cyanidation for a prolonged period under ideal laboratory technique will not dissolve it. Unlike the gangue values, finer grinding has comparatively little effect on reducing the pyrite values of the cyanide tailing.

Before the cyanide residue from the main plant can be floated, it is necessary to destroy the lime alkalinity present, and a good deal of experimental work was carried out before a suitable design of spray tower for conditioning the pulp with S02 gas from the roaster was worked out. This is fully described. Following recirculation of the pulp through this tower to give a pH of about 6.0 (see page 305), it is floated in Fagergren cells

160


ROASTING

161

using reagent 301, copper sulphate, and pine oil. The concentrates are thickened, filtered on a drum filter provided with a flapper (see page 120), and fed by belt conveyor at 16 to 17 per cent moisture to the Edwards roaster. On the basis of a mill feed of 1200 tons per day, about 23 tons per day of concentrate carrying 26 to 27 per cent sulphur is. roasted. . The installation includes two standard 70-rabble roasters, though only one unit is currently used. A 25-hp. motor drives the complete mechanism of each roaster. Special insulation was provided on the side walls and arch, and by using air-cooled rabbles a close control of temperature is possible without the use of any outside fuel. A cooling hearth is incorporated in each of "the Lake Shore roasters. This arrangement is made by providing a lOH-in. drop in the bottom of the roaster, which completely stops back mixing of the charge between the cooling hearth and the roaster proper. In the original roaster, eight pairs of rabbles out of the thirty-five are used for the cooling hearth. In the second roaster, the number was decreased to six pairs. The wet concentrate is a sticky, puttylike substance which adheres to almost any dry surface. The back mixing action of the rabbles ensures that there is always a quantity of dry, disintegrated charge at the feed end of the hearth. The fresh feed entering the furnace falls into this bed of dry, dusty material and picks up a coating of dust. This action can be compared with the baker's use of flour to prevent bread dough sticking to his bread pans. The dust-coated or dust-lubricated concentrate will not adhere to dry surfaces, nor will it ball up. The rapid turnover and thorough mixing given by the rabbles distribute the dust-lubricated material over the first three or four bays of the roaster, where it can be dried by the heat of the charge and the gases. The first pair of rabbles are double armed to reduce the amount of wet feed which piles up under the chute between passes of the rabble. The moisture is driven off, and the filter cake shreds break up into more or less equidimensional lumps. On further application of heat, the trapped water inside these smaller lumps turns into steam. The pressure set up inside the lumps causes them to burst.

An over-all hearth slope of about 7~-in. to the foot is provided. However, it is stated that at no point in the furnace can the charge be said really to flow. During the elimination of the first atom of sulphur, indicated by the blue flame, the charge appears to be slightly lighter and fluffier. Even at this point, the angle of repose is over 30 deg. Elsewhere, the angle of repose is nearer 45 deg. These angles of repose compare with 35 deg, which is the angle for a rock dump. This absence of flow makes it possible to work with a deeper bed of charge and enables the rabbles to do a more efficient job of turning the charge each revolution. The flocculent nature of the charge undoubtedly contributes materially to reducing the dusting and dust losses to the low figure obtained.

162


The cooled calcine is discharged from the roaster by means of a classifier rake mechanism. The rake draws the calcine out well clear of the discharge port in the roaster before dropping it into the preliming agitator. Operations at Lake Shore have demonstrated the vital importance of this temperature control over the initial stage of the roast. The charge must be held at a low temperature for a sufficiently long period during the blue-flame stage. (The temperature has reached 900°F. at the fifth port and rises slowly to a max...imum of 1150° at the eighteenth port. Exit gases at 500 to 650°F. contain about 2.7 per cent S02.) The ore is roasted to hematite rather than magnetite-the latter condition resulting in a higher cyanide consumption. ARSENOPYRITE

The roasting of ores containing arsenopyrite presents greater difficulties than straight pyritic ores mainly because of the tendency toward the formation of insoluble arsenites and arsenates, which have a detrimental effect on gold recovery.' In consideration of roasting procedure, it is quite generally agreed that reducing conditions should be maintained in the early stages of the roast to ensure elimination of the arsenic in the arsenious state. Provided this has been satisfactorily accomplished, the finishing stages of the roast can be done under active oxidizing conditions. It is not enough that reducing conditions alone be maintained during the period of arsenic elimination, as it is' also necessary to maintain movement of the charge and a good flow of gas over or through the charge to carry off the arsenious oxide as it is produced. A second important consideration in roasting procedure is that of the time-temperature sequence. For Beattie concentrate it proved best to hold the temperature about 900°F. during the period of arsenic elimination followed by a rise to 1300°F. before completion of the roast.

Consolidated Beattie Gold Mines Ltd. This 2000-ton per day plant is operating on an arsenical gold ore in the Duparquet area of Quebec, Canada. The operation consists of straight flotation, followed by roasting and cyaniding of the concentrates. The roasting plant, which is one of the finest installations in Canada, has paid for itself in increased recovery over that obtained by cyanidation of the concentrate direct. The flotation concentrate is thickened, filtered to 16 per cent moisture, and partially dried (8 to 10 per cent moisture) in a coal-fired Ruggles Cole drier. The discharge is in the form of round pellets, which are broken up by rollers placed on the inclined conveyors. It is delivered to 250-ton bins placed above each of three 25-ft.-diameter by 13 hearth Wedge splitdraft roasters. 1

F. R. Archibald, "Roasting Arsenical Gold Ores and Concentrates."

163

ROASTING

The furnaces are provided with an installation of two hot C~ttrells for taking dust out of the hot gases, the 10 to 13 tons per day of dust-carrying 2.25 per cent arsenic being returned to the third hearth by chain drags and elevators, and two cold Cottrells, which treat the gas after it has been cooled to 250 by addition of cold air (formerly by water sprays) to precipitate the arsenic. There is also a powdered-coal-firing arrangement for adding heat to the roasters because of the low sulphur content (16 per cent) of the concentrate, and until recently heat was added to certain hearths continuously. At the present time supplemental heat is not supplied, and a feed carrying as little as 12 per cent sulphur plus 2 per cent arsenic has been successfully handled when a high tonnage rate (up to 190 tons per day) is maintained. The roasters have excellent insulation. 0

TABLE

20.

ROASTING DATA AT BEATTIE

Item

Daily tonnage ............ . . . . Loss on calcina tion, per cent .. Gold, oz. per ton ............. Sulphur as sulphide ........... Sulphur as sulphate ........... Total sulphur ..... ............ Arsenic ....................... Iron ..........................

Crude ore

Flotation concentrate

1200

180

. . . . . ..

......

0.16

1.11

...... . ... . ..

......

.

2.20 0.35 4.6

......

16.0 2.0 16.7

Calcine

155 14.2 1.30 0.23 1.55 1. 78 0.3 20.0

The roasters are operated on the split-draft principle; i.e., they are providedwith bleeder flues from hearths 7 and 11 and two uptake flues from the first roasting hearths to a balloon flue. The As 2 0 3 , which condenses out of the gases when cooled below 500°F., is a product that is very corrosive and difficult to handle. The product from the Cottrells runs about 77 per cent As as As 20 3, 12 per cent 80 3 and carries about 0.03 oz. Au per ton. As this is not high enough grade to meet present market demands, it is stored in concrete bins. The roaster-plant gases go to a 410 ft. high concrete stack, 18 ft. bottom and 7 ft. top diameter. The calcine is quenched, without liming, and pumped to hydroseparators which are in closed circuit with tube mills, where the calcine is ground to about 90 per cent minus 325 mesh. This overflow is thickened, filtered, repulped in cyanide solution, and cyanided by a conventional flow sheet. A total of 1066 tons of solution running about $5 per ton is precipitated per day. Zinc dust consumed equals 0.006 lb. per ton. The present over-all recovery is 85 per cent, consisting of a flotation recovery of 90 per cent, a roaster recovery of 99 per cent, and a cyanidation extraction of 93 per cent. Table 20 summarizes the work in 1948.

164


The particle-size distribution of Beattie concentrate, as given by Archibald for 1940 practice, is shown in Table 21. Conditions at various points in the roasters, in 1939 when the plant was operating full capacity are shown in Tables 22 and 23. TABLE

21.

PARTICLE-SIZE DATA AT BEATTIE

Micron size

+ 100 mesh + 56 micron-IOO mesh + 40 micron- 56 micron + 20 micron- 40 micron + 10 micron- 20 micron -

10 micron

TABLE

22.

. . . . . .

2 4 6 8 9 10 12 Bleeder 7 ........ ! Bleeder 11........

Gold, oz. per ton

20.6 30.6 7.8 13.4 9.0 18.6

0.64 0.88 0.90 1.15 1.50 0.86

HEARTH TEMPERATURES

Roaster 1

Hearth

Weight, per cent

Roaster 2

Roaster 3

Temp., OF.

Sulphur dioxide, per cent

Temp., OF.


'I'emp., OF.

400 620 840 1050 1150 1230 920 830 780

2.64 2.82 2.90 3.09 3.18 1.98 0.52 1.11 2.6

420 610 800 1060 1210 1300 1060 770 980

2.64 2.40 3.09 4.50 2.60 1.69 0.47 1.97 0.99

460 650 790 1030 1190 1260 900 740 800

TABLE

23.


2.90 2.40 2.82 2.90 2.44 2.23 2.56 0.81

GAS TEMPERATURE AND VOLUME

Item

Temp., OF.

Gas vol., cu. ft. per min.


Multiclones ................... Cold Cottrell outlet ........ ... Stack .........................

545 250 248

32,200 81,600 82,500

1.85 0.86 0.85

Drafts used: Hot Cottrell inlet. '" Hot Cottrell outlet Cold Cottrell inlet Stock draft

0.44 0.45 0.9 1.05

in. in. in. in.

water water water water

ROASTING

165

SULPHOTELLURIDES CRIPPLE CREEK ORE

Golden Cycle Mill. This roasting plant, since closed -down, was equipped with eight standard 70-spindle Edwards roasters, though only four were finally in use. Fuel was supplied to four ports on each furnace in the form of pulverized lignite of 8500 B.t.u. value prepared at a central plant. About 7 tons of fuel was consumed per 24 hr. for a rated furnace capacity of 120 tons of ore, which is equivalent to 117 lb. of coal per ton of ore roasted. The feed carried 3 to 4 per cent moisture, and the roasting temperature varied between 600 and 650°C. -I'he time of contact was close to 57-§ hr. The last 16 spindles operated in the open, and the aircooled calcine after being sprayed with water was discharged by means of a reciprocating drag conveyor to a Cottrell precipitator. This unit resulted in improved extraction, since the fines were known to carry higher values than the ore as a whole. KALGOORLIE

Lake View and Star. The dewatered concentrates are passed over a Merrick weightometer into a 80-ton bin from which they are fed by a ribbon feeder on to the main roaster feed belt. Eight Edwards duplex roasters are installed, and each is capable of handling 20 tons of concentrates per day. They are fed by a constantweight feeder and are heat controlled by means of auxiliary off-take flues placed along the crown of the roaster. These flues draw off hot gases and reduce the temperature at any given point to the desired figure. Ideal roasting is achieved -when the temperature does not exceed 550°C. during the stage when pyrite is being oxidized to pyrrhotite, and this reaction is recognized by the lilac-colored sulphur flame. The air is drawn through the roasters by two fans having a combined capacity of 100,000 cu. ft. per min. The fans discharge to two six-cyclone Van Tongeren dust collectors and thence to a 9 ft. 6 in. diameter chimney stack 170 ft. high. Loss of weight in roasting = 24.6 per cent. FLUOSOLIDS ROASTING

Dorrco FluoSolids Reactor. Figure 49 shows a generalized crosssectional view of the FluoSolids reactor which has recently been developed and is now being applied to the roasting of sulphide ores, to the calcination of various materials, and in general to problems involving reactions between solids and gases at elevated temperature. In a recent paper on this subject delivered before the Canadian Institute

166


of Mining and Metallurgy at the October meeting in Winnepeg, Manitoba, Owen Mathews? states: The term "fluidization" denotes the fundamental principle of carrying out a gassolid reaction in a dense suspension of solids, maintained in a turbulent mass by the upward flow of the gases that effect the reaction. This mass assumes a fluid level and acts as a fluid. The upwardly moving gas stream imparts to this mass a turbulence resembling that of a boiling liquid. The outstanding advantage of this principle, as embodied in the FluoSolids system, is the close degree of temperature control realized and the uniformity of the temperature condition that can readily be maintained throughout the fluidized bed. Under normal conditions there is no appreciable temperature difference between parts of the bed, nor is there any appreciable temperature difference between the gaseous phase and the solid phase.

This is believed to be highly important, because it means that the heat released is immediately distributed throughout the entire bed. Thus there is no observable local overheating at any point or points, resulting in fusion and locking-up of the gold values. Furthermore, heat losses are relatively low, leading to the belief that roasting can be properly carried out on sulfide concentrates containing as little as 12 per cent or even less sulphur and without the use of purchased fuel. Roasting at Cochenour Willans. At Cochenour Willans, the FluoSolids reactor consists of a cylindrical shell built of H-in. steel plate 18 ft. high with an inside shell diameter of 8 ft. 8 in., reduced to an effective inside diameter of 6 ft. 8 in. by a lining of 9-in. firebrick backed with 3 in. of insulating brick. The domed top of the reactor, the hot-air ducts, and the cyclone dust collectors are lined with a castable refractory cement. Appropriate openings are provided, piercing the steel shell and lining for pressure taps, thermocouples, and feed and discharge connections. There are two outlets in the dome top of the reactor, one, 14 in. in diameter, being a gas discharge to the cyclone dust collectors and the other, 8 in. in diameter, being merely an auxiliary stack opening which is used only when the reactor is being preheated prior to being started up. Both are lined with refractory material. The main 14-in. gas line leads to two cyclone dust collectors arranged in series. The stack is 14 in. in diameter, built of :J,B-in. steel plate, and is 125 ft. high. In the base of the reactor shell there is fitted a steel perforated constriction plate. This plate is lined with castable refractory cement and contains 120 cup-shaped orifices in each of which a 3-in. Korundal sphere is seated. These spheres act as distributing valves for the diffusion of the air throughout the bed and as check valves when the reactor is shut down 2 "FluoSolids Roasting of Arseno-pyrite Concentrates at Cochenour Willans Gold Mines, Ltd" by Owen Mathews, mill and roaster superintendent.

167

ROASTING

or, in the case of power failure, to prevent calcine from passing down through the perforations into the conical air chamber below .. Through the conical base and constriction plate an 8-in. pipe is fitted with an external butterfly for emptying the reactor when required. All external piping is of wrought iron. Three quench tanks are provided to receive, respectively, the hot calcine from the reactor, the dust from the first cyclone, and the relatively finer dust from the second cyclone. All the quench tanks discharge to a STAtK

FAN

a

ID:

~~~

CYCLONES

FIG. 49. Dorrco FluoSolids reactor.

common, screened receiving tank, the contents of which are pumped by a 2-in. rubber-lined sand pump to the cyanide plant. The arrangement of equipment directly ahead of the reactors is as follows: Flotation concentrates are pumped first to a small thickener then to a small rotary vacuum filter, with filter cake, at 12 per cent moisture or less, going to the reactor and with thickener overflow and filtrate going either to waste or back to the flotation circuit. The filter cake is fed to a 3-ton steel bin through a ;\i-in. vibrating screen and is introduced into the reactor by a variable speed, Coghill-type feeder, consisting of a vibrating hopper and an 8-in.-diameter by-18-in. long ribbon-type screw con-

168


veyor, suitably designed to act also as a seal to prevent gas escaping from the reactor at this point. The current tonnage of concentrates handled in the reactor is 8 tons per day assaying about 6 oz. gold per ton. Since the reactor has a rated capacity of 15 tons pel' 24 hr. (actually considerably more), it is operated on two shifts only. The reactor can be shut down and started up without difficulty and without any apparent disturbance in metallurgy 01' extraction. Because the Cochenour Willans ore is self-roasting, temperatures are controlled at about llOO°F. by injecting water into the bed." Thermocouple connections at various points in the furnace permit automatic recording of temperature at about 5-min. intervals. Manometers connected above and below the constriction plate and at the top of the furnace indicate the pressure drop between various parts of the reactor. For information on the cyanidation of the calcine see the general mill description in Chap. X'V. The Dorrco Fluobolids reactor is also available in multicompartment design, the top or "feed compartment serving to preheat the ore and effect preliminary decompositions, such as dehydration or the elimination of arsenic, the midclle compartment being the main reaction chamber, while the lower compartment serves to cool the calcine and preheat the incoming gases. The, material passes down from one compartment to the next through "overflow" standpipes, and the countercurrent flow of gases and solids makes it possible to effect considerable heat economy where this is desirable. THE USE OF SALT IN THE ROAST

During the development stages of the roasting process at Lake Shore it was found that commercial cyanide extractions from the laboratory calcines could be obtained only from roasts in which salt had been added to the charge. This information was later confirmed in both the pilot plant and- the commercial plant employing Edwards roasters. The exact nature of the chemical reaction produced by salt, which has such beneficial effects in the cyanidation of the calcine, is not clear. While it can be demonstrated consistently on the Lake Shore concentrates, roasting tests carried out at Lake Shore on concentrates from mines in other camps failed to show any improvement when salt was added to the charge. These outside concentrates produced commercial cyanide extractions when roasted without salt. 3 The high degree of heat conservation in this type of furnace requires this step with ores of more than a certain critical calorific value in order to maintain the temperature at a predetermined figure.

ROASTING

169

Certain observations that were made on the effect of adding salt during the laboratory roasting test work have been borne out in the plant roasting. 1. The cyanide extraction was raised from 70 to 80 per cent (ohtuined without salt) to 90 per cent when salt was added at the rate of 30 lb. per ton roasted. 2. The presence of salt in the charge shortened the total time of roasting required to produce a dead calcine by from. 10 to 20 per cent. This speed-up in the roast is accounted for by the charge commencing to calcine at an earlier stage in the roast and probably at a lower temperature than is the case when salt is omitted. This effect applied equally to those outside concentrates which could be roasted and cyanided successfully without salt. 3. The calcine produced without salt from Lake Shore concentrates was a brown color. In roasts where salt had been added, the calcine' was a rich purple-red shade. A fourth effect was noted at the time of the laboratory roasting work, but the investigation of sizing technique has rendered this effect difficult to interpret. The calcine from salt roasts, when given the normal infrasizing treatment, appeared to be finer than the calcines produced with no salt. The use of salt in roasting a gold-bearing concentrate requires a certain amount of caution. A process which depends on the use of excess N aCl and lime has been described for the extraction of gold from a concentrate by volatilization in a roaster. To avoid loss of gold, the amount of salt added must be held at such a figure that this volatilization does not occur. The addition of 30 lb. N aCl to the Lake Shore concentrate does not cause any measurable volatilization of gold in the plant roasters. ROASTING WITH SODA ASH

Studies by Prof. G. J. MacKay of Queens University, Ontario, Canada, connected with special methods of roasting arsenical ores resulted in the finding that:' After heating arsenical ores or concentrates mixd with soda ash equal to 5 to 10 per cent of their weight in the absence of air for a period of 20 to 60 min. at temperatures between 950 and 1200°F. and then quenching in water, residues very low in gold values could be obtained by eyanidation of the calcine.

It was later found that 94 pel' cent of the gold in a 1.92-oz. concentrate could be extracted by quenching in water alone without the use of any cyanide at all. It is known that gold forms soluble sulpharsenate and polysulphide compounds, and it is believed that the gold was taken into the quench solution through such agency. It could be precipitated from the solution by gentle aeration, particularly in the presence of a trace of manganese salt. Such precipitates assayed from 30 to 40 oz. per ton in gold, and the gold so precipitated was not readily soluble in cyanide solution.

There appeared at the time to be a number of difficulties attending the commercial adaptation of such a process, but as new roasting techniques 4

Archibald, op, cit.

170


are developed, it is probable that further work along these lines will be carried out. GOLD LOSSES IN ROASTING

In general, serious gold loss by volatilization is not experienced in roasting gold ores or concentrates provided that the chloride content of the feed is below a certain critical figure (at Lake Shore this was 30 lb. pel' ton approximately) . Considerable gold loss may be experienced in the case of arsenopyrite ores, however, if the temperature is allowed to rise too rapidly during the fuming off of the arsenic. N. S. Spence of the Department of Metallurgy, TABLE

24.

CHEMICAL ANALYSIS OF FLOTATION CONCENTRATE

Item

Iron ................... Arsenic ................ Sulphur . . . . . . . . . . . . . . . Copper ................ TABLE

25.

Per cent

Item

Per cent

26.85 15.52 19.30 0.20

Antimony. . . . . . . . . . . . . Insoluble .............. True silica .............

0.16 10.3 7.8

CONDENSED RESULTS OF ROASTING TESTS

Roasting temp., "C.

Per cent loss in weight

Per cent loss in gold

412 491 615 700 802

30.7 30.6 30.6 30.8 32.0

0.7 4.5 18.8 28.1 33.7

Queens University, Kingston, Ontario, describes certain tests carried out in a laboratory muffle furnace as follows. 5 Samples for each run weighed 150 grams. A gas-fired muffle was used, which was brought up to the desired temperature before charging the roasting dishes. Every run was done in duplicate. The temperature was recorded every 5 min., and the average calculated. The temperature used was that of the atmosphere of the muffle directly above (% in.) the roasting dishes and was read from a thermocouple pyrometer. Roasting was continued in every test until no trace of SO, could be detected by smell in the air above the dishes. This period of time averaged 95 min. During roasting, the charges were rabbled carefully every 5 min., great care being taken to avoid dust loss. Duplicate assays were run on the roasted product, and knowing the weight and the assay of each dish, the loss was calculated. • Abstracted from E. and M.J., Vol. 137, No.7.

ROASTING

171

CALCINE TREATMENT

The usual method of treating the calcines from the roasting o{ gold ores is to cool in air, quench the moderately cooled material in water, and subject it to a light grind with pebbles or balls to break up agglomerated particles, sintered prills, etc. The ground pulp is then passed over corduroy or other types of blankets to trap any free gold released in the roasting operation, and then cyanided, by either continuous or batch system. The above procedure is frequently modified to suit local conditions, both blanket. concentration and grinding being omitted in certain cases, while in others water and even acid washes followed by filtration before repulping in cyanide solution are resorted to. oJ A number of examples of calcine treatment will be found in Chap. XV, but the following may be considered typical of this operation. Lake Shore Mines, Ltd," Preliminq. As it leaves the roaster the calcine contains less than 0.1 per cent in soluble sulphur but as much as 2.6 to 3.0 per cent soluble sulphur (as sulphate) and a variable amount of chloride, which probably never exceeds 0.15 per cent. The calcine has an acid reaction, and before it can be handled in contact with iron and steel, it must be neutralized. This is carried out in a preliming step in a 7- by 7-ft. turbo-agitator, which oxidizes the ferrous salts and precipitates the corrosive sulphates and chlorides present as hydrates. About 60 lb. lime per ton of calcine is used to bring the solution strength to 1.0 lb. CaO per ton. Aeration. There are two stages of aeration in the treatment of the calcine pulp. The first stage is in the preliming agitator, where the pulp is given a thorough aeration. This aeration oxidizes the ferrous iron in the pulp to the ferric condition, in which form iron has a negligible cyanide consumption and no oxygen consumption. It also serves to saturate the pulp with oxygen before any cyanide is added. The second stage of aeration is in the calcine agitation. At this point it is necessary only to maintain the oxygen content of the pulp. Aeration in the calcine agitators is obtained by blowing compressed ,air into the pulp through perforated pipes wrapped with six layers of canvas. These aerators are made of 8-in. lengths of I-in. pipe, four per agitator. The aerators are located as near the bottom of the tank as possible to take full advantage of the increased solubility of oxygen in water due to pressure and to provide a maximum time of contact between the pulp and the rising column of bubbles. The use of an efficient aerator in the calcine agitators has materially 6

Abstracted from Williamson's paper referred to under "Pyrite." I

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reduced the calcine cyanide consumption. However, the canvas sock aerator requires cleaning or replacement at 9-week intervals. Cyanidation. It will be recalled that the flotation concentrate contains certain free gold values which resisted dissolution during the 60 hr. of main plant cyanidation. These values, after roasting, dissolve rapidly and completely. The same rapid dissolution also applies to the values originally completely occluded in pyrite. The roast changes the pyrite grains into porous friable hematite which allows the cyanide solution to penetrate to and dissolve most of the gold values. Test work has indicated quite clearly that cyanidation of the Lake Shore calcine is virtually complete after as short a treatment as 6 hr. Mter 9 hr. treatment, further time of contact will not lower the cyanide residue assay. In plant-scale operations, the handling of a small tonnage of calcine presented certain problems. At a dilution of 2 to 1, a single 24- by 24-ft. agitator would, theoretically, furnish over 24 hr. agitation. However, short-circuiting would be so serious as to render any such treatment useless. It had been found in the main plant circuit that even six agitators in series did not completely overcome short-circuiting, However, a compromise was made in designing the calcine treatment plant. Four 12by 8-ft. agitators were used in series. This provided some 24 to 36 hr. treatment. Thus, by more than trebling the time of contact known to be necessary for dissolution of the values, it was believed that the short-circuiting due to the use of only four tanks would be overcome. The continuous calcine treatment required the use of a small Oliver filter. IIi view of the small tonnage to be handled, filtering costs per ton of calcine were high. The operation of the small filter was a constant source of inconvenience. The high sulphate content of the calcine pulp gave rise to another problem. As the calcine pulp cooled during the agitation, CaSO precipitated. This formed a cement to bond together a heavy build-up of calcine around the rim of the agitators. The removal of this build-up necessitated shutting down the agitator and chipping off the accretions, which at times were well over a foot thick, In 1945, the continuous process was abandoned in favor of batch agitation. The calcine filter was scrapped, and batch filtration was carried out on one of the main plant filters. The same agitators are used as in the continuous process, but for batch treatment they are arranged as two pairs of tanks. The pulp from the calcine mill is pumped to one of the pair of tanks until they are filled. The flow is then diverted to the other tanks. Cyanide [Aerobrand C~(CNh] is added during the filling period so that, when the batch is filled, the cyanide is approximately at full strength. Additional cyanide is added as required to maintain a cyanide strength equivalent to 0.5 to 0.7 lb. KCN per ton of solution during the treatment. The time of treatment, i.e., from the time the tanks are filled until the pulp is run off to the filters, varies from 12 to 20 hr. but is never less than 12 hr. This allows a sufficient margin of safety at all times. One of the regular main plant filters is used to filter the calcine pulp. The main

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plant pulp is cut off shortly before the calcine batch is finished, and the filter is allowed to work out any solids remaining in its tank. Then the calcine pulp is fed to it. Once batch agitation was instituted, most of the difficulty of a mechanical nature disappeared from the calcine cyanidation. The danger of short-circuiting was removed. All the calcine pulp' received a known time of treatment of sufficient length to ensure dissolution of' all values recoverable by cyunidat.ion. . The trouble previously experienced owing to build-up on the rim of the agitators disappeared. Since batch agitation, there has been no build-up at any point. The precautions taken in cleaning out the residual main plant pulp before using a filter for calcine are necessary to prevent the calcine filter discharge from being diluted by the main plant pulp. It is desirable, for control purposes, to have a reliable assay of each batch of calcine residue. oJ Calcine pulps are considerably more difficult to filter than the normal main plant pulp. The spongy, porous nature of the calcine adds to the difficulty of washing the cake. However, with careful operation it is possible to recover over 99.5 per cent of the dissolved values in a single stage of filtration. For satisfactory filtration, the calcine cake must not exceed ),<1: in. in thickness. Wash water must be used in sufficient quantity to cover the cake completely at all times. Failure to keep the cake covered results in the formation of cracks. Cracks in the filter cake provide a path of negligible resistance and so decrease the amount of wash water passing through the cake. The filters are fitted with five spray pipes. The first three carry barren solution and the last pair fresh water. The calcine treatment is carried out in a solution 0.5 to 0.7 lb. per ton of cyanide as KCN. Little improvement in gold extraction can be gained by the use of higher solution strengths, which do, however, result in higher cyanide consumption.

Rietfontein (T.C.L.).7 The roasted calcine withdrawn from the furnaces is collected in specially designed trucks. These act as cooling and storage bins, as well as a means of conveyance, and are fitted with a special bottom-discharge arrangement. By the use of these trucks the problems of cooling, storing, and dusting were eliminated, as no intermediate transfer of the calcine from truck to bin or bin to agitator was necessary. As each truck is filled with calcine, it is sampled and tested for unroasted calcine. The hot properly roasted calcine is shunted onto a cooling track from which it is subsequently transferred to the leaching section of the plant. About 330 gal. water and 200 lb. of 90 per cent sulphuric acid are added to the agitator, the paddle gear of which is then set in motion. Seven trucks of calcine, equivalent to 2.45 tons, are fed to the agitator, and agitation proceeds for 1 hr., after which the acidity of the solution is tested. The acid-treated pulp is fed by hose from the agitator to two 5by 2-ft. box vacuum filters, each box taking about 22.5 lb. per sq. ft. of filter area. A dry vacuum pump connected to the filtrate receiver is put into opera7

S.A.1VI. and E.J., July 10, 1948.

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tion. Calcine pulp is fed to the filters to within 1 in. of the top of the boxes, and the vacuum valves are then opened to allow filtration to commence. When the original copper sulphate solution has been filtered from the calcine and the cake is surface dry, a water wash of 2 in. is run into the boxes, followed by five washes of 1 in. When full, the vacuum receiver is emptied via a l-in. acid-resisting pump, which delivers the copper-bearing solution to two sand clarifier storage tanks, each 6 by 4 by 5 ft. 3 in. The clarified solution from these tanks is fed at controlled rates to two copperprecipitation boxes, wherein the copper is precipitated from solution on steel scrap. The barren solution is allowed to run to waste. The precipitated or cement copper is cleaned up twice a month and is one of the materials still dispatched overseas for realization, since all attempts to find a local market were unavailing, owing to the association of arsenic with the copper. The cyaniding of the calcine is effected in three mechanical agitators, while the gold-bearing cyanide is separated from the calcine by means of a 3- by 2-ft. Denver rotary filter. The filtrate drawn from the Denver filter is pumped to two sand clarifiers, whence it gravitates to five filiform zincextractor boxes arranged in series. The extractor-box effluent gravitates to a storage sump, from which solution is drawn to supply the sprays on the Denver filter and for "make-up" solution required in the agitators. The zinc-gold slime cleaned up from the extractor boxes is treated with sulphuric acid for the removal of the zinc and subsequently with nitrosulphuric acid to remove any copper which has found its way into the cyanide section of the plant from the acid leaching section. The solution siphoned off after this operation is delivered to a "cyanide" cement precipitation box of three compartments, steel scrap again being the precipitant. The gold slime is then transferred from the acid vat to a filter box in which it is dewatered, The dewatered gold slime is calcined and subsequently fluxed as follows: gold slime (by weight) 58 per cent, borax 20 per cent, sand 12 per cent. The manganese dioxide is 10 per cent. The fluxed gold slime is smelted in a reverberatory furnace and poured into button molds. The buttons are then remelted in a carborundum crucible, granulated in water, and fluxed with 42 per cent (by weight) borax, 33 per cent sand, and 25 per cent sodium nitrate. This flux serves to remove bismuth which is present in the gold. GOLD LOSSES IN CYANIDING CALCINES

An interesting problem confronting metallurgists is how to reduce the gold loss in the residues from the cyanidation of roasted calcines. An extensive investigation into this matter is reported in two recent papers published in Australia: "The Condition of Refractory Gold in Lake View and Star (Kalgoolie) Ore" by N. 1. Haszard and "Roasting and Treatment

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175

of Auriferous Flotation Concentrates" by A. F. B. Norwood. The conclusion reached, as reported in the first of the above papers, was tha't of the three forms in which the gold occurs in Lake View and Star concentrates, viz., (1) as free gold, (2) as gold tellurides, (3) as gold associated with pyrite, the residue losses were almost entirely due to incomplete solution of the. gold associated with the pyrite. Apart from the evidence of microscopic study, roasting separately various sized fractions of the concentrate gave practically identical results, which again points to the extremely fine state of subdivision in which this refractory gold occurs. A series of tests also showed that the presence of calcium salts in the calcine had an adverse effect on cyanide extraction, presumably through the formation of compounds that reduce the porosity of the iron oxides resulting from the decomposition of the pyrite during roasting. In the second paper, which reports a continuation of the same investigation, a rather detailed study was made of the chemical reactions taking place at various stages in the plant roaster by taking samples at various points along the hearth. An analytical method is described for determining elementary sulphur, sulphur as pyrite', and sulphur as pyrrhotite, also the proportions of magnetitie and hematite present in the calcine. Tests were next carried out in which pyrite and pyrrhotite were roasted in sulphur dioxide. Since the reaction is endothermic, the temperature of the grains cannot rise above that of the furnace and the harmful effects of "flash roasting" is avoided. While the cyanide results were little better than the best roasting in air, it is worth noting that the magnetic or "black roast" is quite as amenable to cyanidation as the conventional "red roast."8 As a result of tests carried out at various temperatures, it was concluded that the high cyanidation tailings obtained following the higher temperature roasts are due to recrystallization of the iron oxide which destroys the porous structure induced by roasting and consequently locks up the submicroscopic particles of gold in dense crystals of hematite. Various methods for reducing the rate of combustion in plant furnaces are discussed. Residual values of about 1.8 dwt, per ton appear to be the lower limit with present practice. While laboratory attempts to sulphate roast have not been very successful, the conversion of the calcine to ferric sulphate by strong sulphuric acid and subsequent ignition back to ferric oxide provides a material which will give almost 100 per cent extraction to cyanide. It is the opinion of the author that future improvements in recovery would appear to lie in the direction of devising a method of speeding up the oxidation of pyrites at temperatures below 500°C. The gold can be removed from calcines by volatilization with salt in an B This has been the experience of the Dorr Company in the testing of certain Canadian ores using FluoSolids technique.

176


oxidizing atmosphere, and also by smelting with lead, but there are economic and technical problems involved in both of these schemes. Gold-silver-arsenic Alloys. Tests carried out by the staff of Nepheline Products, Ltd., in Canada in cooperation with F. R. Archibald to determine the solubility of gold-silver-arsenic alloys in the cyanide solution led to the following conclusion: ... arsenic, alloyed with gold-silver alloys, aids rather than hinders dissolution of gold in cyanide solution, and such could be ruled out as a cause of refractory behavior of gold-bearing arsenical ores. The presence of particles of precious metal in cyanide solution residues remains unexplained.

CHAPTER XI

Amalgamation and Bullion Recovery In the present chapter information is given regarding the use of direct amalgamation following stamps or primary mills and the use of barrel asnalgamation· for treating .fig and corduroy concentrates. Precipitation of gold and silver on zinc dust orJon zinc shavings is fully treated, from the clarification and de-aeration of cyanide solutions to the cleanup and melting of the bullion. The llse of aluminum dust, sodium sulphide, and charcoal as precipitants receives attention. AMALGAMATION

Principles. The amalgamation process, which has been widely used in a variety of forms from earliest times, depends upon the wetting and alloying of metallic gold and silver with mercury. Because of its high specific gravity and tendency to be collected by larger masses of liquid mercury or mercury-coated surfaces, this alloy, or amalgam, is readily separated from the lighter constituents of the ore pulp. Richards and Lock' remind us that: In milling, three amalgams of gold may be considered. The first is liquid and when filtered contains only about 0.1 per cent gold at 60°F. The second is solid and represents the combination, in some definite chemical proportion, of gold and mercury. The third form consists of nuggets of gold superficially coated with and cemented together by the two other forms of amalgam. Silver amalgam may be divided into three similar classes.

Mercury also unites with a number of other metals, including copper, lead, tin, zinc, sodium, and potassium and in the case of certain metallic compounds, such as the chloride or sulphide of silver, can under favorable conditions cause them to decompose, with the formation of the chloride or sulphide of mercury and silver amalgam. DIRECT AMALGAMATION

Direct amalgamation, which refers to processes which contact the whole of the ore stream with mercury or mercury-covered plates and which has saved so much of the gold of the world in the past, is now largely obsolete. Modern plants almost invariably employ a concentration step ahead of amalgamation and subject only a relatively small bulk of high-grade concentrate to amalgamation treatment. This scheme greatly simplifies 1

Textbook of Gre Dressing, 3d ed., McGraw-Hill, 1940.

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the cleanup operations and reduces the chances of gold loss through theft and other means. However, a description of some of the old methods may be of interest. Amalgamation at Argonaut, Calif. Mortar-box and plate amalgamation at the Argonaut mine, Mother Lode district, used to save 70 pel' cent of the gold, according to S. E. Woodworth in I.C. 6476, U.S.B. of M., 1931. Amalgumation is practiced both inside and outside the 12-st::Lmp batteries. For inside amnlgamation, straight-sided, copper-faced, wooden chuck blocks are used. These blocks are equipped with half-round iron strips, spaced 2 in. apart, for their entire length. It was found that the half-round strips assisted amalgam to build to a greater thickness than did the smoother block. If the hourly inspection of the chuck blocks indicates that mercury should be added, it is fed in measured quantities from a horn spoon with the incoming ore at the back of the batteries. The chuck blocks are cleaned twice a month or oftener. The total quicksilver fed is recorded and gives a close estimate of the free-gold content of the ore crushed. The pulp passing the battery screen falls upon reverse splash plates, the first of which is 5 by 51 in. in size with an area of 1.77 sq. ft. set at ::L slope of 3 in. pel' ft., and the second is 8 by 51 in. in size with an area of 2.83 sq. ft. and an inclination of 4 in. per ft. There is a drop of 2 in. between these plates. On leaving the reverse splash plates the pulp drops 5 in. to the cast-iron lip of the mortar.. This drop is variable, depending on the height of the battery discharge. From the mortal' lip, the pulp falls 4 in. to an amalgamated apron plate, set at a slope of U:b in. per ft. This plate is 49 by 58 in. in size and has an area of 19.4 sq. ft. An amalgam trap consisting of a wooden box of length equal to the width of the plate and of a uniform depth of 8 in. is attached to and forms a part of the apron frame. The pulp flows from this trap through four 2-in. iron nipples, set in the side of the box, on a plane 4 in. from the bottom. The pulp, issuing from these nipples, drops through a lO-mesh, woven-wire, brass screen to the sluice plate. If a battery screen is punctured, this lO-mesh screen catches the coarse oversize. The total drop from the discharge nipples of the trap to the sluice plate is 6 in. This plate is set at an inclination of U:b in. pel' ft. and is 46,l>:l in. by 16 ft. in size. It has an area of 62.10 sq. ft. The total length of plates pel' battery is about 21 ft., and the total plate area is 86 sq. ft. On the average tonnage this represents about 4.3 sq. ft. of plate area per ton of ore milled per day. As the pulp leaves the sluice plate, it drops into another amalgam trap which is attached to and forms a part of the plate frame. This trap is a wooden box built the entire width of the plate, but unlike the first trap it has a sloping bottom, and the depth of sand adjacent to the plate discharge is but 4 in., whereas the depth along the trap overflow is 6 in. All plates, including the chuck blocks, are made of copper which weighs 5 lb. per sq. ft. and are electroplated with 3 oz. of silver pel' square foot of contact surface.. When the tube mills operate and grind the sands from the classifier, which treats the tailings of the first two rows of vanners, the ground product is amalgamated on eight shaking plates 4 by 5 ft. in size, which have a combined area of 160 sq. ft. These plates are set upon a slope of % in. per ft. and are oscillated ninety times a minute by a simple strap eccentric. Each morning the amalgam on the apron and sluice plates is softened with mercury, rubbed with a rag, cleaned with a rubber squeegee, and dressed with a whisk

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179

broom for the d:iy's operation. The cleaning and conditioning of these plates require about 3 1 :5 hr. The monthly cleanup, including retorting of amalgam and melting of bullion, occupies 5 days. On the first day of the cleanup all sluice plates are cleaned and scraped with wide-faced putty knives, and in addition one battery is dismantled, cleaned, and reassembled. During the second day six apron plates are cleaned and scraped with a scraper made from an old file, the end of which has been flattened, widened, sharpened, and turned at right angles to its length. Two additional batteries are cleaned on this day. On the third day, the remaining six apron plates are scraped and cleaned, and three more batteries are cleaned out. During the fourth day all splash plates are taken to the cleanup room to be steamed and scraped, and, in addition, four batteries are cleaned up. The two remaining batteries are cleaned on the fifth day. The amalgam recovered from dressing the plates eaclr'day is squeezed into a pellet and stored until retorted with the general cleanup amalgam. Residues taken from the batteries are placed in an amalgam barrel, which contains three pieces of stamp stem and which is rotated for about 12 hr. Mercury, amounting to 350 troy oz., is then placed in the barrel, which is again rotated for an additional hour or two. It is then stopped, opened, cleaned, and the pulp run into a storage box under the barrel. From here it is fed by the cleanup man to a power jig, the bed of which forms on a fine wire screen. The amalgam is found in the hutch of the jig with the iron floating on the amalgam, and the sand over the iron. When all of the barrel charge has passed through the jig, the sand is scooped from the hutch, the iron is removed by a magnet and the liquid amalgam is removed through a spigot into an iron dipper. Here it is further cleaned by mechanical agitation and a water jet which removes any foreign matter present. After squeezing the cleaned amalgam in a canvas cloth by hand and removing most of the liquid mercury, additional mercury is removed by further squeezing the soft amalgam in canvas, using a mold and a hydraulic press. The final amalgam from the cleanup, placed in trays, is sealed in the retorts during the afternoon of the fifth day. The retorts are heated, and the quicksilver vola.tilized and condensed during the night. Wood is used for fuel. The next morning, the sixth day, the bullion sponge is removed from the warm retort, placed in graphite pots and, after melting in oil-fired furnaces, is poured in bars which are shipped at once either to the American Smelting and Refining Company at Selby or to the United States mint at San Francisco. The trays into which the amalgam is placed for retorting are painted with a chalk-and-water emulsion which is thoroughly dried before using. This coating assists in a clean removal of the gold sponge after retorting. During 15 years the bullion has averged 821 parts gold, 159 parts silver, and 20 parts base metal. The mercury loss is 0.17 troy oz. per ton of ore milled.

Homestake Amalgamation. Homestake experience favors amalgamation as a cheap method of gold recovery preceding cyanidation. Laboratory tests at mill sizes indicate that upward of 75 per cent of the gold is free. Of this, about 60 per cent is caught by mercury which is fed to the rod mills and ball mills in closed circuit with Clark-Todd amalgamators and DOlT classifiers in the South plant, the only one now operated. The amalgamator is shown in Fig. 50. It provides new means of increasing the catch of amalgam, by causing the pulp to change direction several times, each change resulting in a retarded velocity of flow and affording opportunity for amalgam to build up on the plate. In this appara-

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tus the launders are fitted with amalgamated trays, which facilitate the attachment of amalgam particles. The trays are deep enough to retain a considerable body of amalgam without loss. In the amalgamator the amalgam is not caught in a thin film 01' layer, spread over a large surface of plate; rather, the catch is three dimensional. The amalgam builds upon itself to form bodies of appreciable thickness.

(b)

FIG. 50. The Clark-Todd amalgamator as installed at the Homestake, South Dakota. (a) As set up at the discharge of rod mill; (b) disassemble~ for cleanup.

The amalgamator also removes scrap iron, fragments of the grinding media, and other material at the point of discharge of the mill, thus keeping clean the recovered amalgam. It does all this in much less floor space than did other systems. The chip screen is removable, but the guard screen below it is not. The second launder is lined with a removable amalgamated-copper box. The bottom of the third launder is covered with an amalgamated silver-plated copper plate. This plate is also removable and is taken out and cleaned at more frequent intervals than is desirable for the trays.

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Homestake amalgam yields 43 per cent gold; mercury consumption is oz. troy per ton of ore crushed, the cost of which was (in 1936) 1.25 of the 1.4 cents total cost for amalgamation. Kolar Practice. Formerly, copper plates in front of the stamp mortar boxes and in the tube-mill circuit, also mercury wells with baffle boards below the plates, saved 78 to 89 per cent of the gold. N ow, by blanket; concentration below the stamp mortar boxes and after the tube-mills, and subsequent concentrate treatment, the recovery is approximately the same as with all-amalgamation. In the blanket machines the blanket or carpet is attached to jute cloth or to discarded filter cloth and is washed by spraying every 1 to 2 hr. Tilting tables to whioh blankets are fixed have been adopted at Champion Reef. The treatment of the concentrate for the recovery of gold values varies on the four mines. At Ooregum the concentrate is treated by amalgamation on copper plates; at Mysore by a process of reconcentration and tabling, yielding gold dust containing 90 per cent fine metals which is smelted direct to bullion; and at Nundydroog and Champion Reef by agitating with strong cyanide solution followed by zinc-box precipitation.

7'8

TREATMENT OF GRAVITY CONCENTRATES

Barrel amalgamation is the simplest and most common method of treating the rich concentrates caught by the corduroy, jigs, or other gravity means. A heavy, cast-iron barrel with manhole, revolving slowly, is used. In it are placed a charge of concentrates, steel balls or a rod or two, some water, lime, and mercury. The whole may be run 2 to 12 hr. The pulp is discharged; the amalgam is then caught in riffles, and the fine pulp in boxes or tubs, from which it may be fed slowly into the mill circuit for cyaniding. The amalgam is retorted in the ordinary manner. 2 Grinding concentrates in batches is the usual practice. Small gold particles are released, and consequently the recovery is greater. However, gold particles smaller than 700 mesh are amalgamated with difficulty, probably because they remain suspended in the pulp. The best grinding practice must be determined by trial, so as to release gold but not comminute the released particles. If the gold is coated with rust, grinding a sandy concentrate usually scratches the gold particles so that they will amalgamate. It is sometimes expedient to mix jig and flotation concentrates for amalgamation. Grinding a sandy concentrate with mercury in an alkaline lime or sodium hydroxide solution will usually give good recovery in spite of the popular belief that flotation concentrate will not amalgamate.

It is recommended that two speeds be provided for barrels where grinding is used, for then it is possible to continue with the amalgamation step without removing the grinding balls. 2

"Recovery of Lode Gold in Jigs," by J. M. Hague, E. and M.J., Vol. 141.

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The time of grinding should be considerably reduced where arsenic and other minerals are present that "sicken" the mercury (see below for "Use of Chemicals"). The capacity of grinding barrels is about as follows: TABLE

26.

CAPACITY OF GRINDING BARRELS

Dimensions,

Diameter X Lenotl:

24 by 36 in. 32 by 48 in. 36 by 48 in.

Approximate Charae

400 to 600 lb. 1000 to 1200 Up to 2000

The barrel is usuallyf' ischarged into a hydraulic, vertical-flow separator, which collects the mercury and allows the fine ore and slime particles to pass on to waste or into the mill circuit. A Berdan pan is effective for cleaning mercury or amalgam or for amalgamating corduroy gold. The pan of this type, made by the Mine and Smelter Supply Company, is suspended at an angle from a guarded, gear-driven spindle within a frame of angle iron standing 4% ft. high. The bowl or pan is 24 in. in diameter and runs at 20 r.p.m., taking % hp, Grinding is done by means of an 8-in. steel ball running loose in the bowl. Material as coarse as 7'4: in. can be fed to the pan. A Berdan pan can be fed continuously or intermittently. If the latter method is employed, when the material has been ground, a supply of water will wash out the slime and leave clean amalgam. This pan can be used for the cleanup, for grinding concentrates at a small mine, or for grinding and amalgamating rich ore. 'The Wheeler pan' consists essentially of a cast-iron tub, usually about 5 ft. diameter and 2~2 to 3 ft. deep, carrying a broad annular die ring on the bottom, on which heavy shoes are drag-ged by means of a yoke; this, in turn, is driven by a spindle from bevel gears and a belt-driven countershaft below the pan bottom. Shoes and dies are ordinarily of gray cast iron which wears down with a rough scored surface. White iron and alloy steels are unsuitable because the wearing faces become smooth and polished, with accompanying reduction in capacity. An adjusting screw with locking wheel is provided for adjustment of the height of shoes. Mullers, which carry the shoes, should be attached to the yoke arms by a flexible fitting in the nature of a universal joint; if a rigid joint like that in an amalgamating or cleanup pan is used, the shoes will often chatter, capacity be reduced, and breakage increased. The die ring and shoe circle are sometimes continuous, but ordinary short spaces are left between both the shoes and the die segments. These form channels into which pulp flows and from which the crushing faces are fed. New shoes weigh 75 to 200 lb., and the crushing force is limited to that exerted by their weight when dragged over the die. Compensating weights are sometimes used to keep the crushing force up to normal as the shoes wear. 3

Taggart, Handbook of Mineral Dressing, Sec. 5-132, Wiley, 1945.


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Use of Chemicals. Mercury losses occur through the formation on the fine metallic globules of a tough skin composed of base-metal amalgams or foreign matter. It is also known that acids liberated from decomposing sulphides attack the mercury, but the latest work seems to indicate that the unsaturated surface of fresh fractures react with the mercury to pro. duce minute films that prevent coalescence. The particular combination of reagents found by A. E. Flynn of the N ova Scotia Technical College to be most effective on an arsenopyrite concentrate from Canada was as follows: Arsenious oxide White lead Sodium hydroxide

0.85% weight of charge 1.50 0.70

It is stated that the lead remains in some insoluble form which eVIdently plays the desired part. Some 99 per cent of the mercury was recovered. Other reagents recommended by operators include lime, lye, cyanide, sal ammoniac, litharge, and even soap, depending upon the particular conditions, but their use is generally the result of trial-and-error methods. Amalgamation at Pickle Crowe;' Concentrates from the cones and blankets are stored in a box measuring 3 by 6 by 3 ft. The contents of this box at 80 per cent solids, with 5 lb. lime and 5 lb. sodium cyanide, are emptied daily under normal operating conditions into the amalgam barrel. This is of cast iron with rubber lining. The inside dimensions are 3 by 4 ft. It is belt driven at 20 r.p.m., and at one time a load of 300 lb. of 2-in. balls was used with grinding continued for 16 hr. Weight of pulp discharge is approximately 600 lb. After grinding and agitation the concentrates are approximately 80 per cent minus 200 mesh. Then 30 lb. mercury is added, and agitation continued for another hour. The use of grinding balls was later discontinued. The door of the barrel is then opened, and the pulp and amalgam allowed to flow slowly over an amalgam plate as the barrel continues to revolve. The plate is silver-plated copper, 3 by 4 ft., with built-up sides to avoid splash. It is set at a slope of 3 in. per ft. The pulp running off the plate is caught in a 12- by 8- by 6- in. trap which holds the free mercury and amalgam not caught on the plate. Overflow from this trap passes over a second plate and a second trap and finally to a strip of corduroy blanket. Blank tails are pumped to the primary thickeners. All traps and plates are cleaned daily. The amalgam is spread thinly over the second plate and washed with water to remove steel, etc. The 4

D. C. McLaren, "Pickle Crowe Gold Mines," C.MJ., November, 1944.

184


corduroy blanket is washed in the storage box. The amalgam is separated from free mercury by squeezing it in a canvas cloth and then retorted in a 2-qt. cast-iron retort. The sponge gold from the retort is melted on refining days, and the condensed mercury added to the working stock. Mercury consumption is 0.5 lb. per charge. RECOVERY OF GOLD FROM CYANIDE SOLUTIONS

Clarification. The first essential to effective precipitation is clear solution. Regardless of color, the solution must be bright and sparkling and entirely free from colloidal solids. One of the most important advantages of proper clarification is the avoidance of undue pressure build-up in the precipitate filters. With proper skill and attention it is possible to obtain such clarification with gravity sand filters, with plate-and-frame pressure filters, or even with vacuum-leaf filters. If canvas is the filter medium, either under pressure or under vacuum, it should be stitched around the periphery of the leaf, and the stitched area painted with "P" and "B" or a similar paint. In starting a new or recently cleaned filter unit, irrespective of type, the effluent should be returned, for a few minutes at least, to the unclarified storage. Close attention to these points will ensure a lower tail solution, less consumption of precipitant, and much more satisfactory melting and refining. Precoating of the filter cloths with diatomaceous earth "filter aids" has been found to improve the clarification operation and to increase the useful life of the filter covers. The precoat when properly used prevents penetration of fine slimes into the filter fabric and thereby extends the useful life of the filter cover. It also assists in removing scale-forming substances owing to the enormous surface area exposed in the precoat, so that pressure build-up in the subsequent precipitate filters from these substances is reduced. A precoated leaf can be washed clean of accumulated slimes more easily and quickly, and consequently less labor and time are required for this operation. PRECIPITATION

Five methods of precipitating gold and silver from cyanide solutions have been used: aluminum, charcoal, sodium sulphide, zinc, and electrolytic. The first and third were specially developed for the silver ores of Cobalt, Ontario; charcoal has been used in Australia, with some possibility of revival; and zinc, either as dust or as shavings, has been used from the beginning of the cyanide process and continues to be the stand-

185


ard method used throughout the world. In this section are given the technique of the processes and their application in certain mining centers. PRECIPITATION BY ZINC

Zinc shavings and zinc dust are both used for precipitation of precious metals. Although most new cyanide plants adopt zinc-dust equipment and some old plants change from zinc shavings to zinc dust, zinc shavings probably will be used at many small mines and tailings operations. Zinc dust, however, is generally more effective and satisfactory than zinc shavings and is approximately 5 cents per ton of ore cheaper than the older method. ~ Chemistry of Precipitation. Chemists differ somewhat on the theory of precipitation with zinc as to whether nascent hydrogen, liberated by the action of an alkali cyanide on zinc, has a direct effect in the precipitation, or is only an auxiliary action taking place at the same time. The following facts are basic: Gold and silver are electronegative to zinc in cyanide solutions and should therefore precipitate them. Precipitation takes place only in the presence of free cyanide. Precipitation is always accompanied by the liberation of hydrogen. The alkalinity of the solution is increased during precipitation.

Clennell states that the entire effect of the precipitation of gold may be expressed by the equation KAu(CN)2

+

2KCN

+

Zn

+H

20

=

K 2Zn(CN)4 + Au

+

H

+ KOH

but that the reaction between zinc and cyanide takes place independently and bears no necessary proportion to it. The following equation represents the probable reaction: Zn

+

4KCN

+ 2H20

= K 2Zn(CN)4 + 2KOH

+

H2

For more details see Cyanide Handbook, by J. E. Clennell; The Cyanide Process of Gold Extraction; by James Park; and Momual of Cyanidation, by E. M. Hamilton. PREPARATION OF PREGNANT SOLUTIONS FOR PRECIPITATION

The requisites of effective precipitation of gold and silver from cyanide solutions are briefly as follows, based on notes supplied by the Merrill Company: Much of the advantage of precoating is lost if the filter aid is not properly used. Attempts to "paint on" the precoat as a thick slurry of filter aid have not given good results. Thinly covered areas are not

186


adequately protected, and a precoat which is thicker than necessary wastefully consumes filter aid. Further difficulties are experienced in attempting to replace a precoated leaf back into the clarifier tank without sloughing off part of the precoat and thereby entirely defeating the purpose of the precoating operation. A precoat clarifier developed by the Merrill Company successfully overcomes these difficulties and has gained rather wide usage. The Merrill clarifier has a precoating arrangement which produces automatically a uniform layer of filter aid of the correct thickness on both sides of the vacuum leaf and permits return of the leaf to service without disturbing or damaging the precoat. The precoating is done in a small, one-leaf compartment usually built in at one end of the main clarifier tank and requires only simple auxiliary equipment, the main elements being a small precoating pump, a float-controlled air-solution separator, and a . source of vacuum. De-aeration or Removal of Oxygen. Efficient and complete precipitation of metals from cyanide solutions requires the preliminary removal of dissolved oxygen. The efficacy of the zinc-dust process is due largely to the preliminary removal of dissolved oxygen from the solution andsubsequent prevention of reabsorption of oxygen in the solution. The Crowe vacuum process is the most efficient and widely used method of de-aeration, since the oxygen content can be reduced rapidly from 6.5 to 0.5 milligrams per liter with a vacuum of 22 in. A method used rather generally at one time on the Rand was to pass the gold-bearing solutions through sand clarifiers having at least 2 sq. ft. of area per ton of solution in 24 hr. To the sand clarifiers mentioned was added fine iron and highly pyritic sand. The solution was deprived of much of its oxygen as it percolated through the bed of sand, especially if two clarifiers were placed in series. Chemical Control. For effective precipitation, solutions must contain enough free cyanide to dissolve the requisite amount of zinc and to hold in solution the compounds that are formed when zinc dissolves in alkaline cyanide solutions. This result is frequently best obtained by adding a drip of strong cyanide solution to the zinc emulsion zone when zinc dust is used. Efficient precipitation of gold and silver solutions by means of the Merrill-Crowe process is generally independent of the strength of the solutions in cyanide and alkali. Substantially complete precipitation is obtained in some plants where solutions contain no more than 0.05 lb. of either NaCN or CaO per ton of solution. In cyaniding silver ores, solutions frequently contain as high as 5 lb. N aCN per ton, with protective and total alkali equivalent to several pounds CaO per ton solution.


187

In cyaniclingmost gold ores, the lime consumption is generally dictated by the requirements of effective settling and is almost invariably higher than necessary for the best precipitation. In cyaniding silver ores, a high alkalinity is needed to dissolve the minerals, and lime consumption for this purpose usually exceeds that needed for settling. The principal detrimental effect of high cyanide and alkali is to consume zinc wastefully. An excess of lime in the solutions will sometimes coat the zinc and choke the filters, rendering frequent cleanups necessary. Precipitation of all gold solutions and of some silver solutions is facilitated by the addition of a soluble lead salt to the solution. Either lead nitrate or lead acetate may be used, although £he former is preferable. The amount of the lead salt approximates 10 pel' cent of the weight of zinc dust if this is added to the solution. The dissolved lead salt is added in the form of a continuous drip to the zinc-emulsion cone or mixing tank or may, under certain conditions, be added to the solution entering the clarifying tank but never with the zinc dust. The lead precipitates as a thin metallic film on the zinc, thus creating an active galvanic couple, with usually more rapid and complete precipitation of the gold and a lower zinc consumption. For some silver solutions the. lead salt should always be added before clarification, because in most solutions some of the lead is immediately precipitated as an insoluble basic salt which rapidly clogs the precipitation filters. The successful use of lead salts requires careful supervision because the addition of an excess at any time may coat the zinc with enough lead to retard or even prevent galvanic action. This explains why lead salts are not used in precipitating solutions containing· considerable amounts of silver, copper, or lead, there being sufficient silver or base metal present to form an effective couple with the zinc. As a precipitant activator, the Merrill Company has determined that sodium bisulphite is of practical benefit where insufficient alkaline cyanide is present. Excess alkalinity must be neutralized to about pH 6.6. Onetenth pound sodium bisulphide is required for a ton of cyanide solution. ZINC-DUST PRECIPITATION

For many years precipitation by zinc dust, as by the Merrill-Crewe process, has been recognized as the most efficient and economical method of precipitating gold and silver from cyanide solutions. Embodying precipitate filters of the plate-and-frame type, of the vacuum-leaf type, or of the more recently developed pressure bag-filter type, the process is in use in the great majority of cyanide plants throughout the world (see Figs. 51 and 52).

188


Advantages. As compared with zinc shavings, more uniform and efficient precipitation is obtained because a fresh surface of precipitant is being constantly exposed to the solution. Where a base metal such as copper or alkaline salts such as lime, magnesia hydrates, or aluminates are present in the solution in large quantities, zinc shavings become quickly insulated or polarized, the boxes must be frequently cleaned, and fresh zinc added. Calcium sulphate in particular is always present in the treatment of sulphide ores and rapidly coats zinc shavings, rendering them inert. In the zinc-dust process the time of contact is so short and the flow of solution through the zinc so rapid that this coating is reduced to a minimum. Even though small, unconsumed particles of precipitant may become coated with copper or sulphates, the amount of zinc thus rendered inert is negligible and is constantly replaced by the addition of fresh, active precipitant. A difficulty frequently encountered in the precipitation of solutions containing dissolved oxygen, particularly in operating zinc boxes in relatively cold climates, is the formation of hydrated zinc oxide or so-called white precipitate. This coats and rapidly destroys zinc and, being mixed with calcium sulphate, is insoluble in acid and causes endless trouble in refining. This compound cannot form in the absence of free oxygen and is therefore entirely absent in plants using zinc dust after de-aeration. Another reason why such uniform and efficient precipitation is obtained by zinc dust is that the filter cloths are at all times coated with a layer of fine precipitant and precipitate and no particle of solution can pass through the filter without first eoming into intimate contact, in faet almost molecular contact, with the precipitating agent. It is this extremely fine state of subdivision of the precipitant which renders zinc dust so effieient. For a given weight of metal the effective surface of zine dust exposed is many hundred times that of shavings, and with this large area a very brief contact between the zinc and the solution is suffieient for eomplete precipitation of the metals. Chemical Considerations. To obtain perfect precipitation, each moleeule of metal-bearing solution must be brought into contact with a particle of precipitant, must give up its metal, and immediately thereafter must be removed from contact with other metal-bearing moleeules. This condition cannot be met in a zine box, and the resultant diffusion accounts for the long boxes necessary and the usual incomplete precipitation. In the zinc-dust process, however, this requirement is fulfilled perfectly, as the solution passes through the layer of finely divided precipitant deposited on the surface of the filter cloth.


189

Under the right conditions, nascent hydrogen is freely formed throughout layer of the precipitant, thus creating the reducing condition necessary for the precipitation of the metals. The actual deposition is due to the action of galvanic couples, consisting of hydrogen zinc, gold zinc, silver zinc, lead zinc, and sometimes copper zinc. The precipitation is closely analogous to electrolytic deposition, with the exception that in the zincdust process it is possible to obtain trace barrens because diffusion or mixing of the impoverished solution with the unprecipitated solution is prevented. Polarization of the cathode particles is minimized by the rapid flow of solution which carries the molecular hydrogen along with it. The precipitation of gold from cyanide solutions with zinc requires either enough cyanide or enough caustic alkali or both to attack the metal with the evolution of hydrogen. Also, the zinc must be in such a form that each tiny bubble of nascent hydrogen will make contact with and adhere to a particle of zinc, forming an active couple. Obviously, this requirement is met much better within a layer of zinc powder than upon the coarse filaments of metal in a zinc box. Therefore, a much higher efficiency of the evolved hydrogen is obtained in the Merrill-Crewe process than with zinc shavings. If the solutions contain dissolved oxygen, the first hydrogen generated is wasted in combining with this oxygen; this, of course, involves a corresponding loss in zinc and alkali and is entirely obviated in the zinc-dust precipitation process, which removes all dissolved oxygen from the solution before contacting with the precipitant. Economics. . The foregoing means that less zinc is dissolved per unit of gold precipitated or deposited. Less unconsumed zinc is left (10 to 15 per cent) in the precipitate; therefore melting and refining charges are less. In silver precipitation the unconsumed zinc is less, being only 3 to 5 per cent. Only 0.6 oz. Merillite or zinc dust per ounce silver is used, compared with 2 oz. when zinc shavings are used. Most of the zinc dissolved in cyanide solutions ultimately goes to form a zinc cyanide, and each pound of zinc combines with 3 lb. sodium cyanide. Subsequently, when this solution comes in contact with fresh lime added to the ore during treatment, part of this combined cyanogen is regenerated-probably less than half, but at least 1 lb. cyanide for each pound of zinc dissolved. Hence, any method that reduces the zinc dissolved in the solutions must also be responsible for a material saving in cyanide consumption. The cleaner (less foul) solutions should also result in a higher extraction of metals from an ore. Equipment. When first introduced, the zinc-dust process utilized the Merrill sluicing-clarifying filter for pregnant or gold-bearing solutions, the Crowe vacuum tank, a zinc-dust feeder and the Merrill triangle-shaped

190


plate-and-frame pressure filters in which to collect the iinc-gold-silver precipitate. Late in 1932, the Merrill Company announced a new form of equipment, the simultaneous clarification-precipitation type. This new type is now employed in capacities ranging from 100 to 1500 tons of solution daily. Many installations of this well-known equipment are in use (see Figs. 51 and 52). Clarification and deaeration of the solution are followed by the immediate addition of zinc and precipitation of the metals without rest and without exposing the solution to atmospheric contact. Most cyanide solutions, after clarification, will, upon standing even a short time, throw out suspended colloids, consisting largely of the hydrates of alumina, magnesia, and iron. Although hardly visible to the naked eye, enough of these precipitates frequently form to coat and "insulate" the zinc, increasing the pressure in the filters and seriously interfering with precipitation. This difficulty is minimized and in most cases entirely prevented by simultaneous clarification, de-aeration, and precipitation, which cost 1 to 2 cents per ton of ore treated. In either the bag or the leaf type of filter, a single, liquid-sealed centrifugal pump effects the successive steps of clarification and de-aeration.' The clarifying tank is. kept filled to a constant level with unclarified goldbearing solution, the inflow to the tank being controlled by an automatic float valve. Suspended in this tank are the vacuum clarifying leaves, with outlets connected to a manifold, which in turn is connected to the top of the vertical vacuum tower in which the solution is de-aerated. The filter leaves after washing are immersed in the precoating compartment to which has been added a small amount of the precoat material, kept in agitation by compressed air. After deposition of the precoat layer, the leaf is returned to the clarifying compartment. Inflow of solution to the tower and, therefore, the solution level within the tower are controlled by an automatic float valve. Within the tower the solution passes down over suitable grids, which break up the flow into small streams and films, thus effecting the substantially complete removal of dissolved oxygen. The top of the de-aerating tower is connected with a dry vacuum pump which rn:aintains a high vacuum within the tower and removes the air released from the solution. The clarified, de-aerated solution is withdrawn from the bottom of the de-aerating tower by a single-stage liquid-sealed centrifugal pump, to prevent reentry of air through the pump gland. Where the bag precipitate filters are used, zinc dust is introduced as the solution flows from the pump to the filters. A belt-type zinc feeder, with motor drive, discharges a regulated amount of zinc dust into a mixing cone. A liquid reagent feeder, operated by the same motor drive,

Unclar/lied solulion

VacWiIl ---..;

I I

!

!

! I

:>

~t"' g2 ~

~

H

o

Z ~

8

Crowe de-aeration tower

I

IJ:j

I

I

q

I

H

I I

I

I I

Pre-coated clarifying leaves

I

t"' t"'

o

Barren solution storage

Z

~

~

to!

I

I I

L~~__~_~~_~=_~~ ~~~~

Pre-coating pump

Vacuum pump

Liquid sealed precipitation pump

a

o· ~

~~~~

Barren solution

to!

~

kj

return pump

FIG. 51. Equipment for Merrill-Crowe press-type precipitation unit using zinc dust.

Note precoat clarifying leaves.

...... ......

c:o

f-'

<0 I>:l

onaarifted solution

~

Vacuum Zinc feeder and

o

~

>-<

t:I

~

le~sifier~

o

!Z

r--

~t:I

Crowe de-aeration tower

o

o

Z o

EI

Z

t-'l

;:d

~

>-<

o

Z

Pre-cooting pump

Vacuum pump

Liquid sealed precipitatiCYI pump

Barren solution returnpump

Clean-up pump

Cleon-up press

o

"J

o

;:d

FIG. 52. Equipment for Merrill-Crowe vacuum leaf precipitation unit using zinc dust.

Note precoat clarifying leaves.

EI

Ul


193

supplies the corresponding and uniform feed of lead nitrate solution to the cone, which is connected to the solution supply tank." cthe leadzinc emulsion is withdrawn from the cone and forced into the main solution line by means of a small' motor-driven, liquid-sealed 'centrifugal pump. The clarified, de-aerated solution now containing the propel' amount of precipitant is forced through the submerged bag filters, the precipitate of the metals remaining within the bags and the barren solution flowing over a measuring weir into a storage tank whence it is pumped for reuse. A pressure solenoid switch is provided, which, in the event of a dangerous rise in pressure, automatically cuts out the"'precipitation pump. In most cases, however, operators prefer to control the pressure by a manually operated valve. Cleanup is effected by emptying the precipitation tank, draining, and then disconnecting the bags and removing the inner filters containing the precipitate. An ordinary washing machine has proved to be quite useful in cleaning the precipitate off the bags. The precipitate is dried, fluxed, and melted in the usual way. The inner bags can be burned and added to the precipitate 01' washed and reused. In the vacuum-leaf precipitate filters, the zinc dust and lead solution are similarly added to a mixing agitator, which overflows into the steadyhead tank supplying the vacuum filters. The mixture of solution and precipitant is continuously circulated over the filter leaves, the barren solution being drawn through the filter leaves by a centrifugal pump, which in turn discharges to a suitable barren-solution storage tank, The world's largest precipitation plant is to be found at the Ranclfontein mine, South Africa, where the installation of 10 Merrill-Crowe vacuum filter units has a capacity for handling 20,000 tons of solution daily. A number of other installations throughout the world are mentioned in Chap. XV. CLEANUP FOR ZINC-GOLD-SILVER PRECIPITATE

Zinc Boxes. As the operations in cleaning up the precipitate from cyanidation of gold and silver ores are so well known, little space need 5 In the Loreto mill, Pachuca, Mexico, where a. silver ore is treated by the cyanide process (see Chap. XVI), there have been some notable improvements in the method of emulsifying and adding the zinc dust for precipitation. Early practice was to emulsify the dust in a portion of pregnant solution and to inject this emulsion by means of a triplex pump into the pregnant flowing in the pipe line to presses. The first improvements came many years ago with the substitution of barren solution into which the zinc dust was emulsified and the replacement of the triplex pump with a small multistage centrifugal. The most recent improvement has been to emulsify the zinc dust in water, which greatly increases its precipitating efficiency, and to use a Shriver high-pressure diaphragm pump to inject into the stream of pregnant solution. c.

194

CYANIDNrION AND CONCENTRATION OF ORES

be devoted to them. Practice varies, but the work is relatively simple and causes little trouble. 'Where zinc shavings are used, generally the fine sludge is taken out of the compartments of the zinc box and later mixed with that from washing the long zinc. The sludge may be acid-treated and then washed, dried, roasted, and fluxed before melting, or it may be only dried and fluxed before melting. At some plants the entire contents of the zinc boxes are acid-treated at every cleanup, but this is not advisable because it entails considerable labor, and as only new zinc is added to the boxes, proper precipitation does not start so quickly as when at least half of the cells are filled with old zinc. When zinc shavings are used for precipitation, less than 60 per cent of the gold and not more than 75 per cent of the silver precipitated are recovered at anyone cleanup, the remainder being returned with the old zinc to the boxes. Filter Presses. Cleanup of the filters used in the Merrill-Crewe process is much simpler and quicker. The precipitates from the filters are uniformly high in gold and silver and in many instances are weighed, fluxed, and melted without preliminary treatment. In most plants the bullion is ready for shipment within 8 hr. after cleanup of the filters is begun. On gold ores, the precipitates assay 60 to 90 per cent bullion with as little as 5 to 10 per cent total zinc, which in most plants is melted direct. Some operators prefer to give a muffle roast before melting, and in a few of the larger plants acid treatment is still used. In such plants, precipitates may be pumped from the precipitating tank direct to the acid-treatment tank, without intermediate hanelling. In large plants the precipitate, either with or without acid treatment, may be melted with litharge, and the resultant lead cupeled, the bullion by this method being of much higher grade. In general, the cupellation method is to be recommended only where large amounts of gold bullion are produced. In precipitating silver solutions, particularly when using Merrillite as a" precipitant, the raw precipitate when taken from the filters contains 75 . to 94 per cent pure silver, and this product is, of course, suitable for fluxing and direct melting without preliminary treatment. The moisture in the precipitate is sometimes reduced to 15 or 20 per cent before melting, but this is not necessary, particularly in the large stationary reverberatory furnaces used in the larger silver mills. An important point in favor of this process and one that should appeal particularly to operators of customs works is the ability to clean up and convert into bullion at any time all the precipitated metals in the plant. Furthermore, the uniform ratio of precipitant to bullion and the fact that a complete cleanup is made make it possible to check accurately the bullion against both mill heads and residues and against solution assays, all of which makes the detection of theft a relatively simple matter.

195

AMALGAMATION AND BULLION RECOVERY ZINC-SRAVING PRECIPITATION

The older method of precipitation of gold on zinc shavings is still used in certain districts. One feature of zinc-box precipitation is that the whole operation can be seen at a glance-solution flow, effect of lead salts, whether copper is coming down, formation of zinc white, and generation of hydrogen. Precipitation on shavings is efficient and reasonably low in cost, but it offers chemical and manual problems not arising in precipitation on zinc dust. Copper and all its compounds readily dissolve in cyanide solutions and form an adherent deposit on the zinc. The -unsuspected presence of copper in an ore is revealed by zinc shavings' becoming red. Precipitation on zinc shavings is sometimes hindered or prevented by the formation of so-called "zinc white," the cyanide and hydrate of zinc. It is insoluble in water but soluble in cyanide solutions and acids. Precipitation at Kolar. At Kolar, India, the group of mines is still using zinc shavings for precipitation, the shavings being cut locally; all solutions are clarified. A typical plant has three zinc boxes with six compartments each (Fig. 53), five of which are used. Each compartment has a capacity of 127~ cu. ft. or 1877~ cu. ft. in 15 cells. A total of 620 tons is precipitated in 24 hr. Average solution feed assays 36 grains gold per ton. Of this 92 per cent is precipitated in the first two cells, and gold is rarely found below the fourth cell. Fresh zinc is dipped in a solution of lead acetate. Zinc consumption is 0.112 lb. per ton of ore treated. Most of the zinc boxes are built of concrete. Box compartments are connected by branch pipes to a main sublevel pipe or side launder discharging into a vacuum-filter tank. Wooden plugs in each cell control the flow to the vacuum-filter tanks when cleaning up. The cleanup proceeds along standard lines and is done two or three times per month. Between cleanups the zinc boxes are rarely dressed. Precipitate is treated with sulphuric acid, roasted, fluxed, and smelted. The bullion averages 985 fine. Precipitation on the Rand. Zinc shavings and zinc dust are both used to precipitate gold on the Rand, all new plants using the latter. TABLE

27.

ZINC-DUST AND ZINC-SHAVING PRECIPITATION ON THE RAND

Condition

Zinc shavings

Assay of precipitated solution, dwt. per ton. Cyanide strength to precipitation, per cent.. Alkalinity (lime), per cent. " . Zinc consumption, lb. per ton milled . Solution precipitated per ton treated .

0.015 to 0.02 0.011 to 0.028 0.005 0.14 1.3 to 2.0

Zinc dust

0.015 0.014 0.018 0.05 1.5

to to to to to

0.020 0.016 0.02 0.06 1.8

1./'1"

t"

I"

.

12x2 flatIron",

2-I!2x;-Z flat bar Fixed by bolts -Y8 "d/arn.,countersunk crt top, holes in bars 8~" apcrrt------ __

I-' ~.

0)

4'!-.............. - - - - - - ---..---y;

~

Z H

t:I

:>-

:::lo

z :>-

z

t:I

F.line,

("J

o Z

("J

I?J

Z

-0 ~

..,:>H

o

Z

o

"J

o

~

I?J

tn

Sectional

FIG. 53. Extractor box, Mysore mine, India.

197


De-aeration of pregnant solution for zinc shavings has eliminated the white precipitate of hydrated zinc oxide. Lead salts are added before precipitation. Wartenweiler, in Trans. 112, .Ii .f.i11.E., 1934, summarizes the two methods as follows: During 1933, 1,055,000 lb. zinc- dust and 3,129,000 lb. zinc shavings were consumed. PRECIPITATION BY ALUMINUM DUST

Precipitation of the precious metals from cyanide solution by aluminum differs from the precipitation by zinc in that aluminum does not replace the precious metals in the cyanogen compound. In the case of zinc the reaction may be expressed by the equation

+

+

2NaAg(CN)2 Zn = Na 2Zn(CN)4 2Ag (Park, The Cyanide Process, p. 180, 5th ed.) or NaAg(CN)2

+ 2NaCN + Zn + H 20 =

Na2Zn(CN)4

+ Ag + H + NaOH

(Clennell, The Cyanide Handbook, p. 123, 2d ed.) When aluminum is used, Moldenhauer, who patented this method in 1893, suggested the following equation: 6NaAg(CNh

+ 6NaOH + 2Al

= 6Ag

+ 12NaCN + 2Al(OH)3

the aluminum hydroxide dissolving in an excess of caustic to form sodium aluminate: 2Al(OH)3 + 2NaOH = Na 2Ab04 + 4H 20 Hamilton (Mamtal of Cyanidation, p. 190) suggests that the following may represent more nearly the actual reaction based upon plant observation: 2NaAg(CN)2

+ 4NaOH + 2Al

=

4NaCN

+ 2Ag + Na2Al

20

4

+ 4H

It is seen that the presence of caustic soda is essential when aluminum is used. Furthermore, as a matter of practical operation lime must be absent at the time of precipitation; otherwise the following reaction will take place:

The calcium aluminate so formed would contaminate the silver precipitate and result in a low-grade product extremely difficult to flux and melt into bullion. Mter precipitation, however, when the barren solution is reused in the grinding and agitation circuits in the presence of lime, the aluminum

198


is precipitated as calcium aluminate and removed from the plant with the tailing, caustic soda being formed. At the Nipissing mill where only a small amount of lime was required and where the solution was already high in caustic from -(;1{e preliminary desulphurizing process, no trouble was experienced with the formation of calcium aluminate in the press. In order to overcome the ill effects of lime in aluminum precipitation when treating ores where the use of a fairly high amount of lime is necessary to promote effective settling, Hamilton and Crawford devised a treatment at the Butters Divisadero mine based upon the following reactions: Ca(OH)2 + Na 2C0 3 = CaC0 3 + 2NaOH CaS04 Na2C0 3 = CaC0 3 Na2S04 (Hamilton, Manual of Cyanidation, p. 195)

+

+

It allows the use of all the lime necessary for neutralizing and settlement; it yields a lime-free solution, for precipitation, and incidentally manufactures the caustic soda necessary for that operation. Apparently, aluminum dust is not effective as a precipitant for gold alone, although the gold in solutions which contains 2 oz. silver or more pel' ton is almost completely precipitated. It has proved its advantages in the treatment of certain silver ores containing arsenic and antimony. It has not been widely applied, however. PRECIPITATION BY SODIUM SULPHIDE

At Cobalt, Ontario. The practice of precipitation by sodium sulphide was developed in 1916 at the Nipissing mill to replace aluminum precipitation. The change was necessary because of the changes in economic conditions wrought by the First World War. The sodium-sulphide process involves the precipitation of the silver as silver sulphide, the reduction of the precipitate to metallic silver by a desulphurizing treatment and the melting down of the resultant silver to a fine bullion. At the Nipissing, precipitation was effected in two wooden tanks, 5 by 6 ft., provided with mechanical agitation. In the first tank the clarified solution met a small stream of concentrated sodium sulphide, which threw down the silver sulphide as a fine precipitate. To avoid blinding the canvas, the precipitate was caused to agglomerate by agitating it in a second tank before it was drawn off to the filter press. In practice it was found that 0.06 lb. sodium sulphide (60 per cent strength) was required to precipitate 1 troy oz. silver. Gold was not precipitated at all, nor was copper, if the solution contains 0.15 per cent or more free cyanide.


199

As with aluminum, precipitation by sodium sulphide regenerates all the cyanide combined with the silver in the pregnant solution. The reactions involved are shown in the equation:

At the Nipissing the precipitate was reduced to metallic silver by a modified form of the Denny desulphurizing process. The precipitate of silver sulphide was transferred to a 7- by 5-ft. iron tank provided with a mechanical agitator. Aluminum ingots weighing 500 lb. were thrown in, and caustic soda was added, about 0.03 Ib.]6 per cent N aOH being required for each ounce of silver. With a dilution of 4 to 1 the solution had a strength of about 8 per cent N aOH. The mixture was agitated until the black silver sulphide turned brown. This required about 8 hr., depending upon the temperature of the solution. The caustic solution could not be too hot, as the reaction with the aluminum would then become too violent. This would interfere with the reduction of the silver sulphide, as the large amount of hydrogen given off prevented the actual contact necessary between the sulphide and the aluminum. The desulphurized precipitate was collected in a filter press, washed free of sodium sulphide, and delivered to the refinery. The aluminum ingots remaining were left in the bottom of the tame for the next charge. PRECIPITATION ON CHARCOAL

T.P. 378, U.S.B. of 111.,1927, by John Gross and J. W. Scott, is a most comprehensive publication on the use of charcoal as a precipitant. It lists 93 references to the literature since 1891 and 7 between 1830 and 1890. Although these items are scattered, doubtless charcoal has been most used in Australia, where, at certain times and in certain interior places, zinc has been expensive. That condition no longer exists, for Australia now produces all the zinc shavings and dust required. Gross and Scott briefly review what had been done prior to their research in about 1926 and then detail their many careful experiments. Their findings are as follows: 1. The mechanism of the precipitation involves adsorption accompanied by a chemical change. 2. Precipitation of silver on charcoal from cyanide solution follows the same laws as precipitation of gold, although it is slower; charcoal has less capacity for silver .than for gold. . 3. The limit of charcoal precipitation from cyanide seems to be about 2000 oz. of gold and 1000 oz. of silver per ton of charcoal. 4. Little difference exists among charcoals prepared from different woods. 5. The most important point in the making of charcoal is the heat treatment, either during the making or subsequent thereto.

200


6. To quench charcoal does not improve it. 7. Pulverization finer than 200 mesh does not appreciably add to the efficacy of charcoal. 8. Few substances in the solution appreciably affect precipitation. 9. The adsorbed gold or silver salt is soluble to some degree in boiling water and is especially soluble in hot cyanide. 10. There is a possibility of so changing the adsorbed gold or silver salt on charcoal that the charcoal may be used for further precipitation. 11. Precipitation of gold on charcoal from cyanide is not metallic and has not the chemical properties of the metal. No gold is visible, even when observed under the microscope. 12. Few substances in solution have a bad effect on precipitation of gold or silver on charcoal, but sodium sulphide and free cyanide decrease the rate. 13. There is a loss of cyanide in charcoal precipitation, due to adsorption. 14. Precipitation of gold is effective from low-grade solutions, but silver is slower, and a countercurrent method is proposed. 15. Some regeneration of cyanide is possible from charcoal when sodium sulphide is used as a "fixer." 16. Charcoal could replace zinc when foul solutions cause trouble in precipitation. 17. A small, isolated plant having wood available could employ charcoal in preference to zinc, using three stages with fairly coarse charcoal. 18. The charcoal has to be burned, and to avoid loss by dusting in doing this, it could be impregnated with sodium carbonate. . 19. Charcoal will precipitate gold or silver from a cyanide-ore pulp; the charcoal can then be separated from the pulp by flotation.

With regard to findings 1, 9, and 10 wherein "adsorption" and "adsorbed gold" are mentioned, Gross and Scott credit A. W. Allen for advancing the theory that adsorption without chemical change of the alkaline aurocyanicle was the correct explanation. This was generally accepted, and all known facts seemed to bear him out. Allen's discussion is to be found in Trans. I.M. and M., 1917-1918, in Vol. 18 of M. and C.E., 1918 (now C. and M.E.); in Vol. 106 of E. and M.J., 1918 and in Julian and Smart's Cyaniding of Gold and Silver Ores. The use of charcoal as a precipitant preceding flotation is covered by Australian Patent 15,635, June 21, 1934, issued to V. T. Edquist, manager at the Sons of Gwalia mine, Western Australia. For more recent developments along the lines of charcoal precipitation, the reader is referred to the carbon-cyanidation process described in Chap.

XIV. MELTING PRECIPITATES

Three types of melting furnaces are in general use, all oil-fired. For smaller plants treating gold ores, the tilting furnace with removable graphite crucible is usually preferred. Clay liners are quite generally used to prolong the life of the pot. In larger plants, particularly those treating silver ores, tilting furnaces of the reverberatory type are most satisfactory. Such furnaces are provided with a molded hearth of suitable refractory and may be either single or double, depending on the amount of precipitate to be handled.


201

In the largest mills treating either gold or silver ores, stationary reverberatory furnaces are used. These may be built to operate as a conventional reverberatory furnace, with the usual fusion hearth of firebrick or other suitable refractory or may be used as reverberatory pot fusion furnaces, with lined graphite pots .. This is the furnace generally used in the large mills on the Rand. _ Precipitate containing up to 30 to 40 per cent moisture may be fluxed and melted without drying, or driers, either steam or electric, may be used to reduce the moisture to from 15 to 20 per cent before adding flux and charging to the furnace. Where filter presses are used, blowing with compressed air yields a suitable product for melting without further drying. Fluxes' used in melting cyanide precipitate vary somewhat in different parts of the world; the following mixtures are typical, but in stm:..ting a new plant trial fusions should be made in each case to determine the most suitable mixture. 6 For melting raw precipitate from clean gold solutions the following charge will usually give rapid fusions and clean, fluid slags. 75 lb. precipitate (15 to 20 per cent moisture). 50 lb. borax. 45 lb. manganese dioxide. 25 lb. silica sand. N aTE. Soda bicarbonate or fluorspar may be substituted for part of the borax, and the manganese may be replaced by niter. Silver precipitate, containing 80 per cent silver or better, is readily melted with the following charge: 100 lb. precipitate (15 to 20 per cent moisture). 5 to 10 lb. borax. 3 to 5 lb. soda bicarbonate. 3 to 5 lb. silica sand. For melting calcined gold precipitates the following charge is used in many plants: 100 lb. dry calcine. 25 to 40 lb. silica sand. 40 to 60 lb. borax. 10 lb. soda ash. 5 lb. fluorspar. NOTE. Five to 15 lb. manganese dioxide may be added in some cases.

Treatment at Bibiani Mine," Smelting of both the calcined precipitate and tabled strake concentrate is carried out by the direct crucible method. For plant-control purposes the two plant products are smelted separately in No. 150 Morgan salamander crucibles fitted with fire-clay liners, and the resultant bullion is collected, remelted in No. 60 Morgan salamander crucibles, and cast into bars. The fineness of the bullion obtained is Corduroy strake bullion Cyanide precipitate

910 fine gold 995 Dore 890

Bulletin of the Merrill Company, "The Merrill-Crowe Precipitation Process." "Treatment at the Bibiani Gold Mine, West Africa," Bul, 492, I.M. and M., 21 pp. 6

7

202


The two products are fluxed as shown in Table 28. The smelting is carried out in oil-fired furnaces that were fabricated at Bibiani and are a modification of those in use at the Ashanti Goldfields Corporation, Obuasi, Gold Coast. Crucible life averages 50 pours, while that of the fire-Clay liners is 8 pours. When new fire-clay liners are fitted to the crucibles, the space between the two is carefully filled and rammed with dry-ground plumbago obtained from old discarded crucibles. All worn-out fire-clay liners are broken up and ground to pass an ;s-in. mesh screen. Any large prills of gold found are collected and join the smelt, while the fines are sent for retreatment with the slags. Treatment of Slags. Approximately 350 lb. of slag is produced monthly, and at the end of each monthly cleanup, this is broken down to 1 in. in size, ground to approximately 100 mesh, and passed over the concentrating tables used for dressing the strake concentrate. TABLE

28.

FLUX USED AT BIElANI

Calcined precipitate

Roasted strake concentrate

Item

No. of parts

Item

No. of parts

Strake concentrate ......... Borax ...................... Silica ...................... Manganese dioxide ......... Soda ash ...................

100 40 25 20 Nil

Calcined precipitate . ..... Borax ..................... Silica ...................... Manganese dioxide ......... Soda ash. . . . . . . . . . . . . . . . o.

100 44 27.5 22.0 11.0

The fine free gold concentrate from the slags is deironed by a magnet, fluxed, and smelted with the cyanide precipitate. The slag tailing gravitates to a settling sump, is allowed to settle, and the water run off. This final product is then dried, sampled, assayed, and shipped to England for treatment by the Tavener process. Slags shipped range from 5 to 10 fine oz. gold per month. . Other products-such as caloiner tray chippings, and floor sweepingsare fluxed and smelted separately. All furnaces are hooded, and ducts are led off to a bag-chamber dust extractor. This is cleaned out every 3 01'.4 months, and the dust smelted separately. Precipitate Treatment at Noranda." The precipitate recovered by the pilot unit was low grade, averaging 35 to 50 per cent copper and less than 2 per cent gold. As the amount of this precipitate was small, the necessity of refining it was avoided by charging it into the smelter anode furnace. The same procedure was followed during the first year the 8

"Cyanidation at Noranda," Bui. C.I.M. and M., Vol. 49,1946.

AMALGA MATION AND BULLION RECOVERY

203

cyanide plant was operated, but on account of the larger amount of precipitate involved, it became apparent that, in order to account for the gold satisfactorily, some other method of treatment would have to be developed. Removal of the base metals with acid and melting of the' residue to gold bullion was tried on a small scale. Best results were obtained when a leaching period of 5 hr. was employed and when the equivalent of 2 to 3 lb. sulphuric acid per ton of precipitate was used at a dilution of 1 to 1 and a temperature of 100°0., with periodic additions of manganese dioxide totaling not more than 0.2 lb. per ton of precipitate. This procedure gave base-metal eliminations "as follows (in per cent): copper, 96; zinc, 99.9; lead, 17.9; and iron, 66.6; and it left a residue that was 13 per cent of the original weight of the precipitate. From this residue, gold of a high degree of fineness could be produced. The method in detail was, however, quite long and promised to be expensive. TABLE

29.

FLUX USED AT NORANDA

Ingredient

Soda ash Borax Niter Silica Lime Iron filings

. . . . . .

Original. per cent

Current, per cent

46.2 15.4

7.7

29.4 30.0 21.0

18.4

7.0

12.3

1.4 11.2

It was decided, therefore, that the best procedure would be to melt the precipitate and produce a copper-gold bullion which could be sampled without difficulty and then be charged into the smelter anode furnace. Melting tests with this object in view were then carried out in sillimanitelined crucibles. These tests showed that the flux required a high sodaash content to ensure fluidity of the slag and that sufficient metallic iron had to be used to keep the copper reduced-otherwise the slag carried appreciable amounts of copper and precious metals. They also showed that best results were obtained with approximately 2 lb. of flux per pound of precipitate. Existing practice is to melt the precipitate to bullion once each month in an oil-fired reverberatory furnace having a capacity of 2000 lb. per 24 hr. The bullion then is remelted in a No. 275 crucible furnace and cast into slabs 15 in. wide, 27 in. long, and weighing 160 lb. After sampling by drilling, these slabs are charged into the smelter anode furnace. The slag is crushed, sampled, and also sent to the smelter, where it is charged into the converters. The composition of the original flux and of the one currently in use is given in Table 29.

204


Treatment of Hollinger Precipitate. In the Bul, I.M. and M., March, 1931, (see also M.M., April, 1931), M. B. Scott describes in detail the treatment of Hollinger precipitate for the production of fine gold bars and the recovery of silver. At this date (1947) the process is still being used successfully, with some alterations in method from the original, the principal one being the recovery of silver by the Thurn process. There are three main steps in the system: (1) the treatment with muriatic acid to remove all the soluble base materials; (2) boiling with sulphuric acid, similar but with the necessary modifications because of its nature to the usual acid-parting process, this being possible, since, although the ratio of gold to silver would ordinarily prevent this, the two precious metals exist together not as an alloy but as a finely divided mixture; (3) final washing, drying, and melting of the gold residue, in which TABLE

30.

ANALYSIS OF HOLLINGER PRECIPITATE*

Item

Gold Silver .. " Copper Lead Zinc Iron and alumina Lime (CaO) Silica Sulphur

* The

. . . . . . . . .

Precipitate, per cent

Residue, per cent

35.0

75.85 15.37

7.2

1.4

2.27

9.3

0.15 1.54 0.14 0.32 0.46 0.76

14.6 0.8 11.7

0.7 4.4

above precipitate and residue are not related but may be taken as typical.

sodium acetate is used to remove the last traces of lead, and the recovery of the silver from the parting acid. The minimum fineness permitted for the gold is 995, but it is not uncommon for its fineness to be 998. Silver is precipitatedas chloride, filtered, and reduced to metal, using scrap steel from discarded boiling pans; dried; melted; and cast into anode plates for electrolysis. The final bar silver, which is practically free from gold, is approximately 999 fine. All slags, old crucibles, floor sweepings, and anything that mayor does contain gold values, together with high-grade or specimen ore from the mine, are smelted in a cupola furnace to produce a lead bullion that is cupeled, and the resulting gold bullion is added at the melting of silver for anode plates. In cases where precipitate from outside sources is handled and it is not convenient to treat this for fine-geld production, the muriatic acid step is


205

successfully used, as the residue obtained by this means contains but a minimum of impurities and therefore, after calcination, it is very easily smelted in a small furnace to produce a gold bullion of excellent quality for shipment. . Bullion Parting at the Homestake Mine. The following information on Homestake practice is of some general interest and is taken partly from an article in JJ1. and JJ1., December, 1935, by Nathaniel Herz and partly from notes compiled by the author in June, 1945. The Miller or chlorine process of parting was adopted by the Homestake Mining Company in 1933, following the abandonment of the gold standard in the United States and the subsequent in~rease in the price of gold. Because 60 per cent of Homestake bullion comes from amalgamation and the balance from cyanide precipitate, this company had a choice of three methods by which the combined bullion could be treated, viz., the TABLE

31.

FINENESS OF BULLION AT THE HOMESTAKE

Parts per 1000 recovered by Item

Gold, , Silver, Copper Lead , Iron Selenium,

,

, ,

,

. . . . . .

Amalgamation

Cyanidation

800 to 820 175 to 190 5 to 10 None About >~ None

760 to 780 185 to 200 25 to 35 None Trace

Miller, electrolytic, and acid parting, and they chose the first. The following analysis is representative of the bullions to be treated. The cyanide precipitate is melted in a Monach furnace, and a double pour is made: one-half the slag in one pot and the balance in a second pot. Mter 5 min. the slag is poured, leaving a scull. The bullion is set in a special divider placed in the pot. The accompanying flow sheet (Fig. 54) gives a good description of the parting and refining operation by which all Homestake bullion is now being treated. All refined bullion is shipped to the United States mint. Silver is melted into bars averaging about 9S0, although a process was developed whereby silver of standard purity, 999 fine, was made. Mter the process had ,been developed, new Treasury Department rules made it unnecessary to refine silver to this degree. With ordinary care, gold bars, as shipped, are about 997 fine. "Treatment of Gold Precipitates Containing Copper." A paper with this title by Norman Hedley and J. J. Kress, Trans. 4S, C.I.M. and M.,

206


----.-

Furnace linings Cottrell sludge, crucibles, tubes and covers and t m::Jaus cleanup

Crude bullion

t I'-----l.~. parting

Meltinn furnace --"r=

Borax

I

Chhrme

g Sir furnace

~~-.~

=r.c: cutt ~fl

Refine/gOld bar

701, --

Crude sl7ver chloride

Borax

ead flux litharge borax sand

t

SI7ver chloride furnace

-r

. ". Excess Iron Castmto and slugs slabs

Borax:i SI7ver

t;

.. Filter

-

I

f Lead

2!!9. totwaste

Bullion

H2S04

t

Ferrous solution to waste

Reduced sl7ver DI7ute HCI

Cup..ellation furnace

gold 8s17ver return to bullion melt

Lsoa~vat~Dilute T

Hightrade fines

LOJr: coarse

...~

sodoash -

~

Graded

H20

'lest

=r=

~o waste

Ft7frate

t

Melting furnace

-r=

-s: ---rTo mint

. - - - - - - - -....Blast furnace

:t

=:t:=

Siafl (Pb-Cu-FeJ Granulatedand returned to secondary Remelt with iron mills borings I Lead tUlllon

t

Iron -copper matte shigped to smelter

FIG. 54. Flow sheet of bullion-refining operations at the Homestake, South Dakota.


207

252-258, 1945, describes a method evolved at the Ore Dressing Laboratory of the American Cyanamid Company for treating gold precipitates containing copper. This comprises treating the precipitates with acidified ferric sulphate, followed by precipitation of the dissolved copper and any dissolved gole,., by metallic iron to give a product analyzing 60 pel' cent copper.' In the case of a gold precipitate carrying 10.06 per cent Au; 0.56 per cent Ag and 26.21 pel' cent Cu, a final product carrying 40.32 pel' cent Au, 2.25 pel' cent Ag, and 8.27 pel' cent Cu was made, representing a 75.05 percent reduction in weight. TABLE

32.

GOLD PRECIPITATES

CONTAI~ING

COPPER

Recovery of gold

Per cent

Reagent

Per cent in leached precipitate .......... Per cent precipitated with copper ....... Total .................................

99.92 0.075 99.995

Ferric sulphate ..... Sulphuric acid ...... Iron ................

Lb. per ten of gold ppt.

0.88

1.10 0.37

The advantages of this scheme are listed as 1. The weight and bulk of the precipitate would be reduced and then less flux and fuel would be required for melting. 2. Special reagents would not be required to flux the copper, with the result that less corrosive slags would be formed and the life of the crucibles, would be prolonged accordingly. 3. Large quantities of slags and material carrying various amounts of gold and requiring special treatment would not be formed. 4. Preliminary roasting would not be necessary, and possible losses due to dusting and handling would be avoided. 5. The copper would be recovered in a high-grade product and credit for it and for the contained gold would be obtained at the smelter. GOLD RETAINED IN CIRCUIT

As soon as gold-bearing ore has been crushed or ground, some of the freed metal lodges in corners or behind mill liners and other places. Pulp that is spilled or splashed from launders, tables, and other machines carries gold, which becomes lost temporarily. Pumps, elevator sumps, and boots are other points where precious metal is held up. Gold is also retained in the. mortar boxes, copper plates, and riffles of stamp mills, especially those in which amalgamation is a feature. This applies also to concentrating and classifying equipment. Ball-mill and tubemill liners retain gold-on the Rand as much as 1000 oz. if grinding is done in water and a tenth of this if done in solution. Some mills in Can-

208


ada retain as much as $25,000 in gold until cleaned out. Cyanide solutions that contain gold are absorbed by wood, and the metal is there held. Wood staves from treatment vats have been known to assay $2000 per ton and were burned to save this gold. The precipitation and melting rooms always lose some metal. The dismantling and cleanup of many plants have yielded much gold to the owners or to others who have bought them. This temporary disappearance of gold, particularly in mills, is sometimes embarrassing to assayers and metallurgists, who are at a loss to account for discrepancies. However, after a new plant has been running for several months, a balance is struck, and all gold is accounted for. Gold finds fewer places in which to lodge in the modern plant with its comparatively small amount of equipment and concrete floors which can be washed frequently. Of the gold produced at the Pioneer mine, British Columbia, 4 per cent of the total was recovered from the ball mills and classifiers when the mills were relined.

E·

CHAPTER XII

Plant Control lJ1any factors enter into the control of operations in a cyanide 01' flotation plant. Those considered in this chapter include cyanides used and the effect of im.puriiie« in cyanide solution; the role of oxygen in cyanidation and the aeration, and de-aeration of solutions, pH control fjnd conditioning, ore and soluium sampling, sizing tests, onuomatic recording, and controls. CYANIDE

The cyanides are compounds of the radical cyanogen CN. Their history covers more than two centuries, but that of the simple cyanides dates back about 90 years (see Edward Thorpe, A Dictionary of Applied Chemistry, Vol. 2, pp. 437-475, 1921; also Wilhelm Bertelsmann, Die Technologie del' Cyanverbindllngen,1906). Prior to the introduction of the cyanidation of gold and silver ores in lS90, the world consumption of potassium cyanide was less than 100 tons a year. Now, 'including the cyanide used in casehardening, electroplating, flotation, and fumigation, world consumption exceeds 30,000 tons of the calcium, sodium, and potassium salts. Gold mines of the Rand consumed SOOO tons in 1933. Flotation processes in the United States use about 500 tons a year of the 12,000 tons for the whole country. Currently, throughout the world, approximately half of the cyanide used for are treatment contains 9S per cent and the remainder 49 per cent sodium cyanide equivalent. The principal American and Canadian sources of cyanide are Niagara Falls, N.Y., and Niagara Falls, Ontario. The Roessler and Hasslacher Chemicals Department of E. 1. du Pont de Nemours Company makes the sodium salt. For many years the Cassel Cyanide Company, at one time known as the Cassel Gold Extraction Company, a pioneer in cyanidation, had a cyanide-manufacturing plant at Glasgow, Scotland. This has been superseded by works at Billingham, County Durham, England. The company is a subsidiary of Imperial Chemical Industries, Limited. The well-known Cassel brand contains 97 to 9S per cent NaCN, equivalent to 129 to 130 per cent KCN. Manufacture of Cyanide. Sodium cyanide is a white, deliquescent, crystalline material easily soluble in water. The basic sources are alkalis or alkaline earths, atmospheric nitrogen, and carbon. In the United States it is derived (1) from sodamide which is produced from sodium and ammonia. The sodamide is heated with charcoal, and the resultant soda cyanamid is then heated with an excess of charcoal, 209

210


resulting in the formation of sodium cyanide. This is the Castner-Roessler process, which yields 96 to 98 pel' cent material. In Europe, 90 to 92 pel' cent cyanide is manufactured (2) from destructive distillation of beet-sugar refuse, forming hydrogen cyanide. This is absorbed in caustic soda solution, from which the cyanide is obtained by evaporation. Aero-brand cyanide is manufactured by fusing calcium cyanamid and salt in a continuous electric furnace, from which the product is tapped at regular intervals into a sump outside the furnace. The conversion is an equilibrium reaction, and practically all of the nitrogen is in cyanide form, By propel' cooling of the melt, the equilibrium is "frozen" at the hightemperature equilibrium point. Samples of this cyanide, kept in sealed containers for several years, showed no change in cyanide content. Types of Cyanide and Cyanide Consumption. Of the two types of cyanide used in the cyanidation of precious-metal ores-sodium cyanide and Aero-brand cyanide-sodium cyanide is generally sold in the form of 5-lb. blocks or cakes and is packed in drums holding 200 lb. net. Aero-brand cyanide is calcium cyanide containing 48 to 50 pel' cent of pure N aCN equivalent, the other half consisting chiefly of common salt and lime. It is manufactured in the form of black flakes and packed for shipment in zinc-coated iron drums. Its dark color is due to a small amount of graphitic carbon derived from the principal raw material cyanamid. This brand of cyanide dissolves readily in water, leaving only a slight undissolved residue consisting chiefly of lime and graphitic carbon. This insoluble residue has no effect whatever on the dissolution of gold and silver and will not precipitate precious metals already in solution. It is, therefore, not necessary to remove it. The recommended procedure for introducing cyanide into a cyanide-mill circuit is to provide a tank of sufficient size so that enough cyanide can be dissolved at one time to furnish the mill requirements for at least one shift. In the case of Aero-brand cyanide, a 10 per cent solution is recommended, which will then contain 5 pel' cent equivalent pure NaCN. It is advisable to aerate the solution with finely divided air (atomized) to eliminate the small amount of soluble sulphides that it contains. The addition of X lb. lead acetate or pulverized litharge will speed up the desulphurizing operation. The tank for dissolving the cyanide should be provided with an agitating mechanism. The strength of solution used in the cyanidation of precious-metal ores will vary with the type of ore and the precious-metal content. Cyanide plants treating gold ores, with little or no silver, rarely use a solution containing over 0.05 per cent NaCN equivalent. In the cyanidation of

211

PLANT CONTROL

silver ores a stronger solution is necessary, and it is common practice to maintain a strength of 0.20 to 0.30 per cent NaCN. Mechanical and chemical losses of cyanide increase in direct proportion to the strength of solution; consequently, the cyanide content of solutions should be kept at the lowest strength consistent with maximum extraction. . The chemical consumption of cyanide, for a given ore, depends on the cyanide-consuming constituents present in the ore, the period of treatment and the strength of solutions. The mechanical loss of cyanide depends on the-type of treatment employed. For gold ores the total cyanide consumption will average around 17~ lb. NaCN pyr ton ore. In the case of silver ores the consumption is generally much greater, probably averaging over 2 lb. High-grade silver ores and concentrates may require as much as 10 lb. N aCN equivalent per ton. (The foregoing notes were prepared by S. J. Swainson of the American Cyanamid Company, New York, which manufactures Aero cyanide. Reference may be made here to four papers on cyanides and cyanidation presented to the Electrochemical Society in September, 1931: "Present Status and Uses of Cyanamid Process Cyanide," by G. H. Buchanan; "Cyanides in the Metallurgy of Gold and Silver," by E. M. Hamilton; "Cyanides in Metallurgy," by M. R. Thompson and "Physical and Mechanical Aspects of the Cyanide Process," by A. W. Allen.) THE SOLUTION OF GOLD

"The Physics of Gold-Solution," by H. A. White, in Jour. C.M. and M.S.S.A., July, 1934, is a highly technical dissertation, wholly confined to the surface reaction between metal and cyanide solution. White's experiments were based on 'the hypothesis that the rate of solution of gold in cyanide solutions is mainly dependent upon the presence of oxygen. The chemical reaction between gold and cyanide may be expressed as 4Au

+ 8KCN +

O2

+ 2H20

= 4KAu(CN)2

+ 4KOH

According to this equation, 1 milligram gold requires 0.0406 milligram oxygen, and this corresponds with 5.80 cc solution at 7 milligrams per liter. At an oxygen concentration of 8 milligrams per liter the corresponding KCN strength is 0.01302 per cent or 0.00980 per cent NaCN. It is found, however, that gold dissolves at the maximum rate if the solution contains 0.027 per cent KCN, equal to 0.020 per cent NaCN, and if it is saturated with oxygen. TIllS difference is due to the slower diffusion rate of the cyanide and to incomplete ionization and hydrolysis for which an allowance of 10 per cent must be made. When the cyanide concentration is lower than the optimum, its diffusion rate will be the determining factor in a saturated oxygen solution, but a

212


greater concentration can only hinder rate of oxygen diffusion as well as reduce its solubility. On the other hand, the presence of excess oxygen, either by means of increased pressure 01' in the presence of oxidizing agents which attack the cyanide slowly enough, 'will raise the optimum cyanide concentration arid the maximum rate of gold solution. If less than saturated with oxygen, as is frequently the case in working solutions, the cyanide strength could be correspondingly reduced, and, in any case, the rate of attack on the gold diminished. It is known that the rate of gold solution is increased by contact with zinc, iron, or carbon, and this may be attributed in the last resort to the extension of surface to which the oxygen may diffuse and likewise involves an increase in the optimum cyanide concentration. Four methods of treatment were employed in the experiments-slime treatment by agitation, sand treatment by percolation, gold plates hung in still solution, gold plates hung in moving solution. The optimum cyanide strength was used, and the oxygen concentration was 7 milligrams pel' liter. These tests, with the physics involved, are given in detail. FACTORS IN DISSOLUTION OF GOLD AND SILVER

A careful study of the factors that influence the rate of .dissolution of gold and silver in dilute cyanide solutions was undertaken by George Barsky, S. J. Swainson, and Norman Hedley and published in Trans. 112; AIM.E., 1934. Cyanide Concentration. The first of the series of experiments had to do with the effect of cyanide concentration on the rate of dissolution of gold and silver. In plant practice, the solution strength for gold approximates 0.05 pel' cent NaCN, 01' 1 lb. cyanide per ton solution. Stronger solutions do not seem to hasten the dissolution or improve the extraction, and as the chemical and mechanical loss of cyanide is much higher with strong solutions, obviously it is desirable to hold the solution at the minimum strength consistent with good extraction. The experiments covered the use of pure gold foil, solutions containing up to 0.50 per cent N aCN and a pH of 9 + but without alkali added. The maximum rate of dissolution of pure gold was reached at 0.05 per cent NaCN, corresponding to concentrations used in modern plants. The solubility of oxygen is practically unaffected by the concentration of cyanide. A similar set of experiments was made on pure silver foil, in 0.01 to 0.50 per cent NaCN. The finding was a maximum rate of dissolution in 0.10 pel' cent NaCN. A third lot of experiments was run on gold-silver alloys containing 79.8 per cent silver and 57.5 per cent gold. Sodium cyanide solutions of 0.10

PLANT CONTROL

213

pel' cent were used. Later, these were assayed and were found to contain gold and silver proportional to the composition of the alloys. ' Alkalinity Variations. A study was made of the effect of varying alkalinity on the rate of dissolution of gold in cyanide solutions. All tests were in cyanide solution of 0.10 pel' cent strength" and varying amounts of lime water or sodium hydroxide were added. The rate of dissolution was greatly reduced at high pH values or high concentration of OH ions. As the curves plotted from the results were so different, further experiments were run to reveal this unexpected action of lime. It was found that lime had no appreciable influence on the solubility of oxygen in the cyanide solutions used, so calcium sulphate-and calcium chloride were added. The former had a slight retarding effect on dissolution of the gold, and the other calcium compound had a more pronounced effect, but as it was determined that the reduction in rate of dissolution of gold caused by the addition of lime is due neither to lower solubility of oxygen nor to the presence of calcium ions, apparently both calcium and hydroxyl ions must be present to produce the full effect, as yet unexplained. DISSOLUTION OF GOLD AS A CORROSION PROCESS

In a recent paper of considerable interest, "The Dissolution of Gold in Cyanide Solution" Trans. Electroctiem: Soc., April, 1947, the author, P. F. Thompson, states: Corrosion research shows that the dissolution of a metal is an anodic process associated with the necessary cathodic action and is therefore called electrochemical; it differs from electrorefining and plating only in that it depends on the electric energy generated within the corrosion cell or local couple itself, and not obtained from without as in these technical operations.

Referring further to the surface agents necessary to maintain the required electromotive potential, it is remarked that: These cathodic agents must necessarily be oxidizing substances but limited in this respect in the case of cyanide, since the cyanide ion may itself be oxidized by most oxidizers to free cyanogen or further to cyanate. Fortunately nature has provided such a reagent in the form of dissolved oxygen, which, though of high oxidizing potential, is restricted in its action by the fact of its slight solubility in water under the partial pressure of 0.2 atmosphere. This limitation harmonized its action with the need to use weak solution of cyanide and also with the relatively minute amount of gold to be extracted.

The author also describes a number of interesting experiments carried out to determine the form of dissolution of gold; the effect of aeration on the time-potential curve; the effect of certain films and of lead, silver, and copper ions on the dissolution of gold leaf. The electrochemical inter-

214


pretation of the results obtained lead to conclusions that agree well with, laboratory and mill experience. For detailed discussion of the physical chemistry of the cyanide process, the reader is referred to the following articles: (1) Barsky, Swainson, and Hedley "Dissolution of Gold and Silver in Cyanide Solution," Trans. 112, A.I.M.E., 660-667; (2) Reynolds "Brief Notes on the Cyanide Metallurgy of Gold," C.M.J., Vol. 65, p. 681, October, 1944; (3) Reynolds "ThePhysical Chemistry of Cyanidation," C.M.J., Vol. 66, pp. 525-530, August, 1945. Among other conclusions reached by the author of the last papers from a study of the mechanics of precipitations is that the real function of lead in assisting precipitation in the zinc-dust process is that of a catalyst to increase the reaction rate of precipitation and that the formation of a zinc-lead couple is only incidental. DISSOLUTION OF SILVER

Metallic silver dissolves in cyanide solution according to the same reaction discussed above for the dissolution of gold. The dissolution of silver sulphide in cyanide solutions, on the other hand, is usually regarded as taking place according to the following reaction: Ag2S

+

4NaCN = 2NaAg(CN)2

+

Na 2S

In discussing this reaction, Hamilton in M an1wl of Cyanidation, MeGraw- Hill, says: This, being a reversible reaction, cannot proceed far before reaching equilibrium, unless the product N a 2S is removed out of the sphere of action. The latter, however, happens to be very sensitive to oxidation, so that a change rapidly takes place probably in two directions:

+ +

+ + H 20 =

1. Na 2S NaCN 0 2. 2Na2S 202 H 20

+

=

N a2S20"

+

NaCNS 2NaOH 2N aOH

+

The thiosulphate would tend later to oxidize to sulphate, and perhaps more sulphocyanate would also be formed.

The need for supplying excess oxygen is evident, however, and the reactions also explain why thiocyanates (sulphocyanates) are always present in solution when cyaniding a silver sulphide ore. The use of a lead salt is usually a material aid to the extraction of silver, and while this has generally been explained as due to its reaction with the N a 2S present to produce insoluble lead sulphide, Clennell" suggests that the lead is rather to be regarded as an aid in the attack on the silver, thus: Ag2S 9

+ PbO + 4NaCN

The Cyanide Handbook, 2d ed.

= 2NaAg(CN)2

+ PbS + Na 0 2

PLANT CONTROL

215

Table 91 compiled from U.S. Bureau of Mines Publications shows the principal silver minerals, with their composition and eyanidation and flotation characteristics. It will be noted that the silver chloride, Horn silver, reacts rather differently in solution from the sulphides and does not require the presence . of oxygen for complete solution. Because of the greater actual weight of metal to be dissolved from commercial silver ores as compared with gold ores, the consumption of cyanide due to dissolution alone is no longer the almost negligible factor it is in the case of most gold ores. Hamilton points out that in cyaniding a 400-oz.-per-ton silver concentrate, as much as 32 lb. cyanide (as KCN) would be consumed in dissolving the silver. In such cases special methods of cyanide regeneration may become necessary. OXYGEN IN' MILL SOLUTIONS

Since the presence of oxygen is an essential factor in the extraction of gold and silver-" by cyanide, the question of the amount of oxygen present in various parts of the circuit becomes a matter of considerable importance. Quantitively this may be expressed as milligrams per liter or, more frequently, as per cent of saturation under the prevailing barometric conditions. Altitude-pressure and Standard Saturation Curves. Chart A (Fig. 55) is used to determine barometric pressure at various altitudes; Chart B is used to determine standard saturation values for various temperatures and pressures. In Chart A, altitudes, in feet, are plotted on the horizontal axis and pressures, in millimeters, on the vertical axis. To find the pressure corresponding to a certain altitude, follow the elevation line downward to its intersection with the curve x, then horizontally to the right, and then read the pressure. For example, if the elevation is 6000 ft., the 6000-ft. line is followed to its intersection 0 with the curve x; then the corresponding pressure, 607 mm, is obtained from the right-hand side of the chart. In Chart B, temperatures are plotted on the horizontal axis, and the amount of oxygen, in milligrams per liter of distilled water, is plotted on the vertical axis; various pressure curves also are plotted, as shown. To find the saturation value for a certain temperature and pressure, follow the temperature line upward until the point corresponding to a given pressure is reached, then follow horizontally across to the left-hand side of the chart, and read off the amount, in milligrams, of oxygen per liter of solution. For example, to determine the amount of oxygen in a solution having a temperature of 59°F. at an elevation of 6000 ft.: The pressure is 10

Except in the case of silver chloride.

216

CYANIDATION AND CONCEN'l'RA'J'ION OF ORES

607 mm; at this elevation, as founel in Chart A,' therefore, the 607-mm pressure curve must be used; i.e., it is necessary to interpolate between the 600- and the 650-mm curves. The 59°F. line is followed to its intersection R with the 607-mm curve, then from the left-hand side of the chart is read off 8 milligrams of oxygen per Iiter. The same procedure is used for the various pressures, using the curve corresponding to the particular pressure. For any particular plant, a solubility curve should be plotted based on the altitude at the plant. Altitucle, Feet O

1-15

00000000000000 00000000000000 o 0 0 0 · 0 a 0 0 a 0 a Q. 0 C!~ N I"J t1"l ..rJ r- cJ:) 01"5 '

<,

1514

I

.\ ~13 ,,\'r-..

:"k

~12

.

640 ~ 600
I

'N. I.\. r-..\ ~11 CJ .\ I\. " SOlturOitioncurves for' I~)(, 4-10 o"''1Qjen in oIistilleel WOlter- CIt........ r-.... o .\ r-,1\..' ..., l'.:: ~6L VOlriou Olltitucles ~

9

:SS

li; 7 0-

C

0

-"

R

6<

""~

I

"- <, ~":-
I'....!'...~~ .......

~5

i

<,

5'2.0 .~ ..460 ~

""....... 440 e lS CO

OO::....~~:::::::r-,

"'sO::: c.:::: t:'t:::' ~ ~ e-,

$4

~

'03

:::::: t-::: :::::: ~s::::: :::::--

~~ E2 ChowtB I ~ ~ t'..:: e 1 I-Computcltions bOlsecl on Winkler's Tebles ~ ~ :-....

:Eo

:E 32

50

68

86

o

10

'2.0

30

104

122.

140

156

Ternperutcre, Degrees F. 40

50

60

70

'"

175

194

'2.12.

80

QO

100

Ternpercclure, Degrees C.

FIG. 55. Curves (A) for determining barometric pressure at various altitudes and (B) for determining standard oxygen-saturation values at various temperatures and pressures.

Results of experiments conducted in the South Kalgurli dry-crushing, all-roasting, and cyaniding plant at Kalgoorlie on dissolved oxygen in mill solutions were given by C. W. Brown in C.E. and M.R., September, 1934. Pulp agitation was done in vats with ordinary stirrers, but considerable compressed air was also introduced. The plant stands at an altitude of 1271 ft. above sea level. The average value of dissolved oxygen just prior to filter pressing the slime was 25.05 per cent saturation, and the average of many determinations in which the agitators were sampled was 28.0 per cent, with a range of 4.2 to 57.5 per cent. As discussed in the next section, a number of dissolved and suspended substances in cyanide solutions tend to reduce the oxygen content, among

PLANT CONTROL

217

which may be mentioned sulphide sulphur, ferrous iron, metallic iron, and organic material. It is noted that, when the cyanide solution is i~ contact with ore the oxygen content is lowered, whereas when it leaves the ore at any point it starts at once to dissolve oxygen. . AERATION

In general, simple exposure to the atmosphere replenishes the lack of ox..'Ygen, but from a practical standpoint appreciable air must be forced into the. pulp undergoing treatment to maintain a satisfactory oxygen content. Chemical oxidizers have been tried, but they are both expensive and in the long run less satisfactory than ail', \vhich is invariably introduced in the agitation step (see Chap. VIII). Aeration or oxygenation of sand, slime, and cyanide solutions has been practiced during the several decades of cyanidation-purposely and incidentally-but not until recently has it been given the careful study that it deserves. Every time a pulp or solution is stirred or transferred, it absorbs some oxygen, but special methods or devices have been developed to entrain air in pulps and solutions. Aeration in the Dorr agitator is obtained from the compressed air used for circulating the pulp through the revolving central lift column and also from the atmosphere when the elevated pulp is redistributed over the pulp surface in a series of small streams from the distributing launders. Additional compressed-air jets are sometimes attached to the revolving arms for greater aeration.. In the Pachuca agitator compressed air used for circulating pulp through the stationary lift pipe is the only source of oxygen. The Devereux and other types of mesh-propeller agitators rely on the vortex created to entrain aIr., The Turbo- and Wallace-type agitators, both of which thoroughly-incorporate air in the pulp through the action of their impellers, are used in several of the Kirkland Lake mills for agitation in small tanks and are also used in the top of and neal' the surface of Dorr agitators to increase normal aeration. Aerating or Oxygenating Cyanide Solutions. A process for aerating or oxygenating cyanide solutions was announced by T. K. Prentice. in the Jam. C.M. and 111BB.A., February, 1934 (see Fig. 56). It imme'':'' diately attracted attention, and the article was reprinted in partby United States and Australian technical journals. The process received practical plant trials at the Nciursemine on the Rand before it was made public, These were mainly on sand which is leached at the mine. Gold extraction was slightly higher when aerated solutions were used, and consumption of cyanide was a third less than in regular treatment. The oxygen content of solutions at the Nourse mine for dissolving gold ranged from

218


27~ to 57~ milligrams per liter and averaged 4 milligrams per liter. At one time it fell to 1 milligram per liter, and a series of high residues resulted. Six milligrams per liter is considered desirable. The oxygen content of circuit solutions in 16 plants on the Rand averaged 4.5 milligrams per liter. In brief, the oxygenating process is as follows, with reference to Fig. 56, which is the patented plant-scale equipment: Cyanide solution is drawn from a stock tank and pumped into the drum or cylinder shown, first passing through the pipe with ~-in. holes to form a spray. Ail' at 100-lb. pressure is generated by the motor-driven (3-hp.) compressor atop the cylinder. The aerated solution leaves at the rate of 2 tons per min., the balanced float valve shown, attached to the discharge pipe, regulating the level of solution in the cylinder. The solution is milk'}' white, owing to the disseminated air, but it clears in a few minutes. At this stage it may

S"pipe, Untreated sctuiion 400g ct.l.. per rmri)

Trecded solufion Sorf'efy release valves _ de/i'/ery presSLfre approx. . ",IOOlb.persq.m., AII' compressor, ,"', Non refurn vetlves,_:..'.... ~2. pipe ' ....,

"~~"

-

/\ )!

/'/

I

5"pipe

, ft ~ . " 'c/ 2"dretin connecf/on I Manhole

I

FICl. 56. Cylinder for aerating cyanide solutions.

carry 6 milligrams of oxygen per liter and retains within 2 milligrams of this amount for 22 hr., which is long enough for the solution to be effective during the first stage of leaching. IMPURITIES IN CYANIDE SOLUTIONS

The following note is taken from Rand Assay Practice, 1932, edited by James Moil' and G. H. Stanley: That cyanide solutions in works practice do not remain pure in a chemical sense is to be expected, having in view their contact during treatment with the many contaminants found in gold or silver ore amenable to cyanidation. Impurities cause a chemical cyanide loss and at times detrimentally affect the extraction of the valuable metals sought. It is fortunate that in the practical application of the process, impurities do not usually accumulate to a prohibitive degree; this is to a large extent due to dilution of the solutions by the replacement by fresh water of the moisture leaving the plant with the residue. With regard to visual signs of impurities, ferro cyanide colors solutions a brownish yellow of varying intensity. Other impurities are generally insufficient to create a characteristic color.

PLANT CONTROL

219

Impurities have their principal source in and are traceable to the ore constituents, to contamination underground and to secondary reactions in the treatrnent at the surface. On the Witwatersrand the oxidation of the pyrite content of the banket reef and wall rock is the chief source. The first place, in order of importance, may, therefore, be given to iron and sulphur in their many combinations and to reactions occurring during exposure in stopes through contact with water (containing dissolved oxygen) so freely applied in the course of dust-preventive measures. Pyrite and marcasite (FeS,) and pyrrhotite (Fe7Ss), where present, are converted (but to a relatively small extent) into soluble ferrous sulphate (FeSO,), ferric sulphate, Fe, (SO,)" and free sulphuric acid (H,SO,,), while colloidal sulphur may be set free. These, in their turn, to avoid excessive corrosion of iron and steel equipment underground, are precipitated by the addition of neutralizing lime as hydrated oxides of iron, both in the ferric (Fe,O,,· H,O) and ferrous (Fe( OHO ),) state, varying according to the degree of completeness of such neutralization and oxidation. The hydroxide in the ferrous state is particularly soluble in a cyanide solution. When these find their way to the cyanide-treatment plant, either with the ore or through the medium of mine make-up water, ferro cyanide (Na,Fe(CN),) and thiocyanate (NaCNS) are formed. This tendency to form acid ferrous salts and thus to destroy the oxygen necessary for gold solution is also latent in the ore undergoing cyanide treatment. As a measure of protection, an alkali, such as lime, is therefore provided and is available throughout; also corrective oxidizing treatment is applied. An oxidation product of the sulphur, sodium thiosulphate (Na,S,03) plays a part. It is often present in the first effluent solution from sand treatment, and its general effect is, by decomposition in passing through the zinc-precipitation boxes, to form an insulating film of sulphide on the zinc which lowers precipitating efficiency. Complete oxidation of this compound to a sulphate appears difficult to attain in practice. Sodium sulphocyanide (NaCNS) is present in practically all solutions in relatively small quantities. It has not been definitely proved to have a deleterious effect on gold extraction. The alkaline sulphide, sodium sulphide (Na,S), resulting from the reaction between the cyanide and ferrous sulphide (FeS) and generally supposed to have a retarding effect on gold and silver' dissolution, is not often observed in solutions on the Witwatersrand, a fact which is due undoubtedly to its precipitation as zinc sulphide (ZnS) by the sodium zincocyanide [Na,Zn(CN),] present in all solutions where zinc isused as a precipitant and also to its oxidation to thiosulphate. In silver extraction its incidence is more pronounced, as the silver itself is often in direct combination with sulphide, as Ag,S in the mineral argentite, pyrargyrite (Ag3SbS,,), and in proustite (Ag"AsS a ) . Lead reagents, such as lead acetate, lead nitrate, or lead oxide, are generally used as a safeguard, acting as precipitants of the sulphide. Resulting from the abrasion and fracture of steel and iron in ore crushing and grinding, metallic iron is found in all mill pulps. Oxidation of this takes place to a certain extent through dissolved oxygen in the water and solutions employed and the aeration of the sand and slime incidental to the treatment process. Any ferrous oxide thus formed is attacked by cyanide solution and is a cyanicide, since the resultant ferrocyanide is practically useless as a gold solvent. With the use of zinc as the precious-metal precipitant, various reactions between this and cyanide take place, the principal compounds being zinc hydrate (ZN(OH),), sodium zincocyanide (Na,Zn( CN ),), and sodium zinc ferrocyanides (Na2ZnFe( CN),)' and (Na,Zn,Fe,(CN)!2l. Sodium zincocyanide is a solvent of gold, as shown by Julian and Smart. The amount of zinc dissolved is considerable, but its retention in solution is not cumulative, as it is precipitated by reaction with the sulphide con-

220


stituents of the are in the ordinary course of treatment and by the ferrocyanide. Sodium cyanide is regenerated in the same reaction, the cyanide loss, therefore, being much less than at first would be expected. Calcium is introduced in the form of lime (CaO) for the purpose of providing a neutralizing agent. Its use results in the formation of calcium carbonate (CaCO,) and of calcium sulphate (CaSO.,). On the Witwatersrand, the treated mine water used as water supply probably introduces the greater portion of the CaSO" content of the solutions. With favorable temperature or saturation conditions it crystallizes out over the entire plant, including the interior of pipes, and may become troublesome. As a physical obstacle and an insulator of zinc, it may be considered objectionable. The use of sufficiently clean water is the best preventive. Sodium carbonate is sometimes used to remove it as precipitated calcium carbonate. Magnesia is introduced to a small extent from underground sources, and finds its way into the solutions as magnesium sulphate (MgSO,) and magnesium carbonate (MgCO,). Gelatinous silica is often found in cyanide solutions. Its effect in practice is more physical than chemical, in clogging filters and extractor boxes. Its source is the action of acid mine waters on the ore constituents. Silica may also be introduced in the form of calcium silicate as an impurity in lime. Organic matter is a common source of impurities in cyanide solutions, its reducing effect being notorious. It is usually regarded as having its origin in mine timber, sewage and sacking, coming from underground and from the surface in the form of vegetal matter and sewage contamination of water used in milling. In the selfdecomposition of cyanide solution, organic compounds such as formates are formed. Prevention is the soundest remedy. Failing this, oxidation by means of chlorine oxidizers has proved efficacious where these can be applied directly or in a separate circuit before cyanide treatment. When strong oxidizers are used on cyanide solutions, free cyanide will be lost by conversion to cyanate. In cyaniding ores containing copper minerals, it is found that the carbonate, oxide, and sulphate minerals particularly are attacked by cyanide with avidity, causing a heavy cyanide consumption by the formation of cuprosocyanide [KCu(CN)2]. In practice this is minimized by the use of extremely weak solutions. Provided that the copper content of the are is not excessive, it is found that copper does not accumulate in the solution, as it is constantly being precipitated by sodium sulphide. ~

CONDITIONING ORE PULP

Conditioning of an ore pulp is understood to mean its preparation for subsequent treatment by cyanidation or by flotation. It is an important step, and its importance is gaining recognition. Conditioning is more effectively done in so-called contact tanks than in ordinary agitation tanks. Generally, a conditioning tank is a small agitator into which the pulp is pumped just after lime, cyanide, or flotation reagents have been added to it, the chief purpose being a thorough mixing. The period of contact is determined in the laboratory and depends upon the nature of the ore, the physical condition of the pulp, and the amount and kind of reagents. The size of the conditioning tank is dependent upon the period of contact and the tonnage to be treated. Broadly speaking, any preparatory machine may be spoken of as a "conditioner." Grinding

PLANT CONTROL

221.

is merely conditioning the ore for cyanidation proper. If fine grinding is essential, then ball mills, tube mills, and classifiers become important conditioners. The thickening of pulp may be termed conditioning for subsequent cyaniding operations.· Air lifts in agitators are likewise conditioners. This idea that the practical purpose of every machine is to condition the ore for the subsequent machine was advanced by L. E. Djingheuzian of the Lake Shore mines, Ontario. Cyanidation. Conditioning operations at the Dome plant in Ontario comprise grinding in water to produce a pulp 83 per cent through 200 mesh which is passed over corduroy. Tailing from the corduroy tables, after the addition of lime, is aerated in four 14::' by 42-ft. Pachuca tanks. The pulp is then treated with cyanide and agitated in a further series of Pachuca tanks. . At N oranda an aerating step precedes each section of the primary roughers in the flotation plant and also on the pyrite recleaning circuit. The machines employed for this purpose are described as aerating classifiers and serve the double purpose of aerating the pulp and classifying into an overflow product for further processing and an underflow for return to the grinding circuit. They consist of circular tanks about 15 ft. deep and of diameters varying from 9 to 16 ft., having a plurality of radial air inlet pipes, four rubber air-lift pipes, and a slow-moving rake mechanism at the bottom. Flotation. The conditioning of ore pulp before flotation is of importance in treating gold and silver ores as well as base-metal ores. Lack of uniform results in a flotation mill can often be attributed to one or all of the following conditions: (1) lack of proper time for chemical and physical reaction; (2) incompletemixing of reagents; (3) fluctuation in the mill feed, causing unequal pulp flow and lack of uniformity in reagent content; or (4) excessive consumption of reagents due to improper conditionmg. The primary purpose of a conditioner, in flotation, is to mix the reagents thoroughly with the pulp before flotation and allow time for the chemical" and physical action to take place. These objectives, of mixing and of completing the chemical and physical reactions, can be attained with some reagents by adding them ahead of the grinding unit. Excessive oxygen beaten into the pulp during grinding precludes the use of certain reagents because oxidation makes them ineffective. Other reagents, because of their frothing properties, affect classification and cannot be used in the grinding units. By using a conditioning tank in the circuit, however, the operator has accurate control of the time of agitation, which can be varied for different ores and reagents. Complete mixing of the reagents is also assured.

222


In addition to mixing the reagents uniformly with the pulp, the conditioning tank serves as a source of uniform feed to the flotation cells, acting as an equalizing tank or stabilizer. Without this tank in the circuit, the flow of feed from the grinding circuit is likely to be uneven and the reagent content irregular; thereby interfering with proper operation of the flotation cells. A uniform feed naturally improves the operating efficiency of a flotation unit. WEIGHING ORE

Weighing of ore by automatic machines is the usual practice of large mills. At small plants an occasional cal' of ore is weighed and an average factor applied to all. Devices for weighing large quantities of ore at a constant rate have always been considered expensive, but such quantities can be weighed with low error. Among apparatus for this purpose are the Blake-Dennison, Hardinge constant-weight, and Merrick weightometer. SAMPLING ORE AND SOLUTION

Proper and accurate sampling is very essential to control of ore treatment. Until a plant has been running for some time, much sampling should be done between the mine and the tailings discharge. Later some of these may be eliminated, or at least certain operations such as screen analyses in certain stages may be dispensed with. Sampling may be divided into three different types according to whether broken ore, pulp, or solution is being sampled. The most reliable sample is that obt~ined continuously. Various machines are made for cutting a. continuoussample from an ore or pulp stream, whereas a drip-wire arrangement is usually used for solutions. If the sample cannot be taken continuously, some method of taking regular cuts is very often used. This may be done by machine or hand. Grab sampling is practiced but of course is not to be recommended for general use. For checking strengths of solutions and during test work grab sampling is often done and is useful. Some sampling practices at different plants are given in the following ex-. amples: Hollinger. Table 33 by E. L. Longmore and M. E. Williams (C.M.J., September, 1935) shows the routine samples taken at the Hollinger, the purpose, and the sampling interval. The Hollinger flow sheet, given in Chap. XV, will aid in interpreting this list, which represents substantially the system still in use. Wright-Hargreaves. The sampling methods at the Wright-Hatgreaves, Kirkland Lake, Ontario, were described by Malcolm Black (Bul. C.I.M. and M., September, 1935). Black notes the importance of accu-

223

PLANT CONTROL

f TABLE

33.

ROUTINE SAMPLES AT THE HOLLINGER

Product sampled

Sampling interval

Assay-ton charge

Hourly 4 times per shift at irregular intervals 2 hr. 2 hr. 2 hr. Hourly

1

Purpose

Mill feed.................... Screen test Classifier overflow; grinding Assay and screen circuit test Table head . Table concentrate . Table tuil . . Concentrate classifier overflow Concentrate agitator head . Concentrate agitator tail .

Assay Assay Assay Screen test and assay Assay Assay and screen test . Assay . Assay

Dorr agitator head Dorr agitator taiL.: C.C.D. tail (solution and pulp) . Pachuca agitator head . Pachuca agitator tail. . C.C.D. No.3 tanks (last) thickener overflow grab sample . C.C.D. sump water grab . Gold solution . Press tails . Press tails, continuous Sample . Repulped primary filter cake (pulp and solution) . Primary filter cake solution. Secondary filter cake (pulp and solution) Secondary filter cake . Secondary filter-cake filtrate (grab sample) . Graphite taken from top of primary thickener 31 (grab sample) . Press-tail drip (drip from gold press~) . Continuous sample . Barren solution by-passes continuous sample during cutting in of gold presses ...

5 2 1 5

5

2 hr.

5 5

2 hr. 2 hr.

5 5

Assay Assay Assay

2 hr. 2 hr. 2 hr.

10 10 10

Assay Assay Assay

2 hr.

10 10 3 to 5 A.T. (Each shift) 750 cc

2 hr.

Color test

Once each shift Every changeover of tanks Hourly

Assay

Each shift

20

2 hr. 2 hr.

10 10

2 hr.

20

Assay Assay Assay and screen test Per cent moisture Assay

Hourly 24 hr.

10

1

Assay Assay

Once a week

10

Assay

At end of cut in

10


rate measurement of mill tonnage. The weightometer in the belt conveyor is checked periodically and is generally correct to within one-half of 1 per cent. Moisture samples are taken every hour. These are placed in a jar having a tight-fitting cover so that drying does not take place before the moisture determination is made. The tonnage delivered to each ball mill is measured by two factors, a revolution counter and the weight of a section of the belt load. The cutter used to remove this consists of two parallel plates attached to and rigidly held by a cross bar. The distance between the plates is equal to one-fifth of a revolution of the head pulley. In taking the belt-weight sample, the conveyor is stopped, the cutter placed down on the belt, and the ore between the plates carefully brushed off into a pan and weighed. This is done every hour. The sample is then passed through a Jones sampler until about 7~ lb. is left. Determination of moisture is made before, the ore has a chance to dry in the air. At the end of each -day the composite sample is again cut down and sent to the assay office. A unique system for sampling various pulps consists of several airoperated cutters, all controlled by one master controller. The master controller is made up of a timing mechanism driven by a Telechron motor, a Geco sampler, a four-way valve, and an air header. The four-way valve is operated by the Geco sampler. The two pressure ports of the four-way valve are connected to a two-partition header. The cylinders operating the cutter are connected to the header by X-in. copper tubing, Wedgeshaped cutters are operated by a cylinder and piston, being made to cut the pulp stream every 15 min. Black also notes the need of a truly representative sample of press heads and tails. The common fault of a drip wire in the whole stream is that the rate of drip is not proportional to the rate of flow. The apparatus devised by Black is illustrated in Fig. 57 and described as follows in E. . and M.J., November, 1944. Sampling Pregnant-Solution (Press Heads). The press-tails solution going to storage discharges into a spill box (il in drawing), thence passing through a bottom pipe into the barren-solution storage tank. The solution will stand in the spill box at a level depending on the volume of flow, i.e., the volume precipitated. On the side of box A is a smaller screen B connected by pipe to the former so that the solutions stands in. it at the same level as in box A. (The screen in this box se!jves to remove putty which is formed by the residual oils and greases acting with the lime.) A swivel-jointed pipe C, tapped into the side of screen box B, draws off the presstails solution in a volume of flow that depends on the height of the solution in the spill box, this volume thus being proportional to the volume precipitated. (In calibrating, pipe C was raised or lowered until the head in spill box A was such as to fill automatic flush tank D in a given time when precipitating a given tonnage.) Automatic flush tank D, next in line, will thus fill from pipe C and be discharged with a frequency that is dependent on the head in the spill box A and so on the volume

225

PLANT CONTROL

precipituted. The press heads are sampled with the same frequency, l1S will be shown, so that the volume of the sample will likewise be propcrtionats to' the volume precipitated. Sampling of the press heads is effected as follows: A portion of the press head, or pregnant solution, flows continuously into nonuutomatic flush tank F, and l1S the tank fills, the sample cup shown fills Iikewise. Each time that the automnt.ic tank D discharges, l1 vacuum is created in leg E' l1S well l1S in sample bottle G, owing to the piping hookup shown. This puts the sampIing nipple attached to the float also under suction, and when the float has dropped sufficiently as tank F empties, the nipple enters the sampling cup and the contents are sucked over into the jar. To prevent continuous discharge of this tank a vent Rodon which float is suspended Float bu.ayed by etectrtc bulbs Pregnant solution

Nut maybe adjustedto suit size of sample desired Rod housingmade/ Vent to stop continuous of ~"pipe discharge by breaking vacuum _Welded / _ ress tails Rubber tube II

B

Screen box

H

Sampling nipple

'-----I'I---:r-"-II-.......? F

c

Nonautomatic flush tank

t

70 anypaint such as gold solution storage

Vent


o

Automatic flush tank

FIG. 57. Apparatus used at Wright-Hargreaves, Ontario, Canada, for sampling cyanide solution. pipe is provided at I. When its end is uncovered, the vacuum in leg E and the sample bottle G is broken and the tank F is ready for the next cycle. The sampling nipple is >s-in. pipe, but this is too large for satisf'actory operation. Therefore, the end Wl1S closed with a little metal by an acetylene torch and >16-in. hole was drilled. This apparutus Wl1S built around an already existing spill box and storage tank. There is no reason, however, why it cannot be adapted to almost any condition so long l1S the method employed includes a tank so placed that the head varies with the volume precipitated.

Sampling Barren Solution (Press Tails). The sample bottles for press tails are kept in a safe so that, if by carelessness the operator allows the press tails to run high, he cannot destroy the evidence. This safe consists of a metal box containing three Winchester bottles. The discharge pipe from the siphon passes through the box, and the sample is drawn off within

226


the box. A small "drip catcher" operated from the outside is used to shift the flow to the various bottles. General. In taking samples of pulp containing cyanide solution, it is the practice at some plants to add a little potassium permanganate solution to the sample container to arrest the solvent action of the cyanide. SIZING AND MINERAL DISTRIBUTION

Wartenweiler, previously cited, made the following concise statement regarding grading tests: "This simple subject is important, particularly where extraction is dependent on degree of comminution and where largescale milling technique must be largely governed by a measurement of particle size." With regard to sieve tests, F. C. Bond and W. L. Maxson, in Trans. 112, A.I.l\I.E., 1934, consider that more information is contained in a screen analysis than is ordinarily recognized and that need is increasing for methods of making this additional information easily available. As more comprehensive studies of crushing and grinding are made, and as the possibilities of decreased expense and increased recovery through a closer control of grinding are explored, methods of interpreting screen analyses assume an importance that they have not had heretofore. The information generally desired includes the size distribution of the material passing the finest screen, the presence of a hard-grinding fraction, the position of natural grain sizes, the presence of sundry materials with markedly different grinding characteristics, the presence and amount of finely divided material (such as clay) which is merely unlocked in grinding, and the total surface area of the ground product. In the United States and in a number of other countries, testing sieves of the W. S. Tyler Company, Cleveland, Ohio, are the standard of metallurgists and technical institutions. This firm manufactures wire clothto the standard scale. Many industries have established 200-mesh cloth as the minimum in screen sizing, and as the U.S. Bureau of Standards has standardized the 200-mesh sieve made from 0.0021-in. wire, having an opening of 0.0029 in. or 0.074 mm, this sieve has been adopted as the basis of the Tyler standard screen scale. When it is necessary to carry an analysis finer than 200 mesh, as is done at Kirkland Lake, Ontario, for example, sieves can be procured as fine as 325 mesh. When discussing pulp sizes finer than 200 mesh, many millmen refer to them as so many microns. One micron is 0.001 mm or approximately H5,OOO in. For the relationship between particle size in microns and Tyler mesh and also a comparison between the Tyler and 1 M.M. series, see Appendix

A. For recommended test procedures in connection with sizing analyses, see Chap. II.

PLANT CONTROL

227

In "Milling Investigations into the Ore as Occurring at the -Lake Shore Mine", May, 1933, to January, 193G, it is stated: The most important and most expensive item in milling is crushing and grinding. The over-all cost of 55 cents at Lake Shore for crushing, conveying, grinding, and classifying is eighteen times that of the next most important item, agitation. Extraction is, of course, directly proportional to the liberation of values. The higher price of gold has meant finer grinding for a new balance between costs and extraction, remembering, of course, that thickening and filtering and difficulty of washing mount directly as the fineness of grind.

FIG. 58. The infrasizer-a pneumatic device for separating a sample of ore into a number of closely sized fractions for analytical and microscopic study purposes.

Because of the extremely fine grinding necessary to liberate the gold values, closer control than that obtained by screens was found to be necessary, and Professor Haultain's "Infrasizer" was successfully adopted as a plant-control and research device. A brief description of the instrument is found in "Fine Grinding Investigations at Lake Shore Mines" (Trans. 42, C.I.M. and M.), which incidentally is probably one of the most detailed studies to be found on the various aspects of particle size control in plant operation. The Infrasizer, TIllS device, which was invented and developed by Prof. H. E. T. Haultain of the University of Toronto, is shown in Fig. 58. In brief, it may be said that infrasizing is an air elutriation process carried on in a series of stainless-steel tubes whose diameters increase by the square root of 2.

228


Thus, the critical ail' velocity in each successive tube will be one-half of that in the immediately preceding tube. Each tube collects a fraction consisting of particles of equal settling rates in ail'. The settling rates of each successive fraction will be one-half the settling rate of the immediately preceding fraction. Assuming a material consisting of perfect spheres, all of one specific gravity, the size range collected in each successive tube will decrease by the square root of 2. However, in the actual sizing of ground products the particles will not be spherical or even cube-shaped but will vary from the extremes of flat lamellae and flakes to needlelike particles, with a fail' percentage of particles of almost equidimensional shape. Not only will the particle shape depart from the ideal condition, but in addition there will probably be minerals of different specific gravity present in any given ore. Such heavier 01' lighter minerals will be sized according to their settling rates. With regard to the meaning of particle size in measuring the fineness of a product, Dr. P. C. Carmen!' says: "Particle size of a nonspherical particle is not a term with a definite meaning unless it refers to a definite property of the particle, e.g., the diame'ter of a sphere with the same volume or with the same specific surface, 'etc. Only for a sphere are all these diameters identical. In the methods of measuring TABLE

34.

THEORETICAL SIZE OF MINERALS IN LAKE SHORE ORE

Relative sizes of heavy minerals, microns

Nominal micron size (gangue)

40 to 28 to 20 to 14 to 10 to o to

56 40 28 20 14 10

Pyrite

32 23 16 11~~

8 0

to to to to to to

Tellurides

46 32 23 16 11~~

8

to to to to 6~2 to 0 to

26 19 13 9H

38 26 19 13 9~~

6H

Gold

to to to to to o to

20 14 10 7 5

28 20 14 10 7 5

particle size, there are measured, respectively, a sieve aperture, a rate of free fall, and some arbitrary microscopic diameter, and the first step is to interpret these in terms of 'particle size.'" Dr. Carmen proceeds to point out the difficulties .and fallacies arising when an attempt is made to interpret the measurements obtained with a microscope in terms of particle size. To avoid confusion in the use of terms, a nominal micron size has been worked out for each infrasizer fraction. This nominal micron size may be defined as the "size" of irregularly shaped particles, of any specific gravity, which have the same settling rate in a column of ail' as do glass spheres of that diameter. As the specific gravity of glass (2.6) is almost the same as that of most siliceous ore (e.g., Lake Shore, 2.7), this nominal size is applicable. In these investigations, all sizing results obtained with the Haultain infrasizer are reported in terms of the above nominal micron size. However, for reference, Table 34 has been drawn up, showing the nominal micron size of pyrite, tellurides, and gold referred to perfect spheres of pyrite, tellurides, and gold, respectively. 11 Carmen, P. C., "The Size and Surface of Fine Powders," Jour. C.M. and M.S. S.A., Vol. 39, p. 268.

PLANT CONTROL

229

The infrasizer has three main uses at Lake 8hore:1 2 (1) for sizing analysis, (2) for gold-sulphur-size analyses) and (3) for preparing samples for panning. For sizing only, 100-gram samples are used and the separation takes 1 hr. 20 min. For the gold-sulphur-size analyses, between 400 and 3000 grams must be sized, depending on the grade of the sample. This. is done in 400- and SOO-gram lots, taking 3~ and 6 hr., respectively. These gold-sulphur-size analyses form the basis for interpreting all the experimental results. From the percentage weight, gold and sulphur assays of

FIG. 59. The superpanner-a laboratory wet-concentrating device for separating an ore sample into a stratified band where the position of each mineral is a function of its specific gravity.

the various, fractions, the gold and sulphur content, and percentage size distribution, are calculated. The superpanner is another device worked out by Professor Haultain for the accurate panning of small samples (see Fig. 59). Professor Haultain's comments on this machine are as follows: This is a mechanized sichertrog or gold-panning horn. The pan is trough-shaped, about 30 in. long by 10 in. wide. The radius of curvature of the bottom diminishes from the lower end to the upper end. A peculiar-shaped cam imparts a bumping motion to the pan, which tends to progress the material toward the upper end. A controllable side motion combined with the wash water washes the surface particles 12 "Milling Investigations into the Ore as Occurring at the Lake Shore Mine," p. 390.

230


to the lower end. There are seven adjustments readily made while in operation. It is a very sensitive apparatus, but as now developed, the technique 01 operation is quickly acquired. It is essentially a batch gravity concentrator. It will give valuable results on as small a quantity of material as 1 gram or will treat much larger amounts. Aided by a binocular microscope or the pyrex-glass method and the microscope, it will readily isolate and identify tellurides when occurring in the proportion of 1 part in 10 million. It will give a clean pyrite product or a clean tailings product. It will separate pyrite from arsenopyrite. It will make separations of extremely fine material down to 15 microns. It has been in the process of incubation for a great many years and was brought in as a useful tool when W. E. Johnston was doing his work on tellurides and was used most effectively by him, and its success is due very largely to his persistence and patience in the development stage.

The superpanner is used at Lake Shore for the following purposes: 1. To determine accurately the assay of the pyrite in the cyanided mill products. 2. To determine the quantity and nature of the undissolved free values in a cyanide tail. 3. To analyze completely original ore samples, high-grade flotation concentrates, etc., separating the sample into the following parts: (a) free metallic gold, (b) free gold tellurides, (c) free galena and altaite, (d) pyrite, and (e) gangue minerals (see "The Form and Distribution of the Gold in the Lake Shore Mill Heads," Sec. IV, p. 294). 4. For preparing high-grade concentrates of the gold-bearing minerals for briquetting, polishing, and microscopic examination. The panner may be used at all times for rough preliminary tests to determine the presence of free values and their nature or to determine an approximate pyrite assay, without sizing the panner feed. However, for all work on which metallurgical calculations are to be based, it is absolutely essential that the samples to be panned should be sized. It is the use in combination of the infrasizer and the superpanuer that has made it possible to analyze completely any mill product. Otherwise, if an unsized material is panned, at best only the coarse pyrite is obtained, and often fine gold or tellurides are trapped with it, giving an erroneously high pyrite assay. The superpanner was particularly useful at the Lake Shore because it was realized early in the test work that the gold in the pyrite was not likely to be recovered by straight eyunidation and that the amount of gold tied up in the sulphide would have to be estimated in order to interpret test results correctly. The panner is by far the quickest and best known means of preparing clean pyrite free from gold and tellurides and gangue. The following is the procedure used to determine the gold content in the pyrite in a cyanide tail: A gold-sulphur infrasizer analysis is first made on the cyanide tail. From this, the size distribution of the sulphur (and pyrite) can be calculated. As the ore contains only 1),2 to 2 per cent pyrite, a flotation concentrate is made in the laboratory to save time in the panning. (NOTE. It has been proved that the assay of the pyrite in the flotation concentrate and flotation tail from a cyanide tail are the same assay per size, so that fact that the laboratory would make only an 80 to 85 per cent pyrite recovery will not affect the assay of the pyrite per size.) Generally, 12,000 to 15,000 grams of cyanide tail is used as feed to the cells to produce 200 to 250 grams of pyrite. If this flotation concentrate is known to contain much free values,

231

PLANT CONTROL

time will be saved by thoroughly cyaniding it before sizing and panning. From each sized fraction, one or more clean pyrite samples are made for gold and sulphur assaying. An average superpanner operator can easily make pyrite concentrates which are 90 to 98 per cent pure. These assays are then corrected to correspond to 100 per cent pyrite. All the plus 14-micron fractions can be readily panned; the 10 to 14 can sometimes be satisfactorily panned. However, by plotting many Lake Shore pyrite results on a graph, a curve for the pyrite assays by fractions has been determined. Thus it is possible to determine graphically the assay of the pyrite in the o to 10 and 10 to 14 (if necessary). All these figures have been checked by calculations made in the routine mill work.

CHEMICAL AND PHYSICAL CONTROL OF SOLUTIONS oJ

In both cyanidation and flotation practice it is necessary to keep a close check on the working solutions, since the concentration of various soluble salts derived from the ore or supplied as a protective measure to the system is usually critical. Alkalinity. Because of the need in most instances for maintaining a definite protective lime alkalinity in cyanide circuits in order to prevent loss of cyanide as HCN by hydrolysis and/or reaction with atmospheric CO 2, the usual titration methods give effective control (see Appendix). In some instances a soap titration'! giving a measure of the degree of hardness in terms of calcium units has been found to be a better yardstick of control than the customary acid titration. Ammonium Sulphate. McLachlan, Ames, and Morton found that at N oranda the presence of sulphates, either added as (NH 4)2S04 or present in the barren solution as CaS04, increases the free settling rate as compared with use of lime water. They also found that a well-aerated pulp settles faster than a poorly aerated one and that the combined effect was beneficial to cyanidation. They consider that the sulphate radical acts as a buffer against the oxygen demand of the pyrrhotite and pyrite. Other authorities state that lime can interfere with the extraction of gold owing to the formation of insoluble cyanogen or other coatings and that the use of ammonium chloride or soda ash will improve extraction in certain cases. There are also instances where, owing to the presence of various cyanicides in the ore,it is necessary to carry solution alkalinity below the range of convenient titration using ordinary indicators, for instance, at Morro Velho in Brazil. In such cases the use of pH measurements becomes of special importance. Some mills also make it a practice to test regularly the pH of the raw water used for make-up. In the case' of flotation, pH control, whether by indicators or recording meters, is very general practice (see Chap. II for principles involved). 13 J. E. Williamson, "A Rapid Determination For Calcium," E. and M.J., March, 1949, p. 75.

232


pH control is applied mainly to pulps in which certain reagents known as regulators are used. These reagents are capable of changing the pH concentration of the circuit and thus influence other factors involved in flotation. Their ability to depress or float selected minerals is closely associated with the degree of alkalinity (or acidity) of the medium. Frequent pH tests should be made. The use of excess lime can be avoided by proper pH control, and it may be pointed out that the use of too much lime not only wastes this material but may have a harmful effect upon the operation. In particular, lime tends to inhibit the flotation of metallic gold. The positive influence of hydrogen-ion concentration likewise holds true in the use of regulators other than lime; moreover, the degree of active alkalinity or pH exerts a characteristic influence throughout the processing not only on the regulators but also on the other reagents used. Hydrogen-ion determinations should be performed on the individual raw ores, because these substances vary widely in their natural reaction, and a record of such characteristics will be of value to the operating engineer. Furthermore, pH determinations are of value in the control of the water used in flotation. Differences in pH always will be found in the raw waters, and these changes often cause variations in the operating results of the flotation machine which cannot be explained otherwise. Alkalinity Control at Sub-Nigel, Successful cyanidation of SubNigel ore, containing about 1 per cent pyrrhotite, requires pre-aeration with close and constant control of pH between 9.6 and 10.0. As neither glass nor antimony electrodes of the usual industrial pH meters can be used directly in pulps containing abrasive solids or cyanide, it was necessary to devise some method of continuously measuring pH which was sufficiently simple and foolproof to be used by the ordinary operator. It was found by experiment that the agitation in the Pachuca tank would continuously remove the cake from a leaf filter suspended in the pulp, so that clear solution could be withdrawn almost indefinitely without interruption. The apparatus developed consists of a I5-in.-square filter leaf made of X-in. pipe and 7E"-in. screen covered with I5-oz. filter cloth. The filter leaf is suspended in the pulp from the rim of the Pachuca tank and is connected to the Merrill-Crowe vacuum through a small solution receiver mounted at the top of the Pachuca, As pH varies indirectly with temperature, the bottom discharge from the solution receiver is carried down through the pulp to prevent change of temperature from the filter to the electrode assembly and pH meter, located under the Pachuca tank. . The electrodes are wiped with a soft cloth once every two or three days to remove any deposit that may have formed; other than this no attention

PLANT CONTROL

233

is needed, and over a period of more than a year no maintenanco has been required. The filtering rate through the leaf filter is 160 to 200 cc pel' min. of clear solution which is sufficient to displace the solution in the electro~le assembly every 2 min. A complete description with diagram and comment on some of the difficulties encountered is given in a paper by H. E. Cross presented at the May meeting of Chemical, Metallurgical and Mining Society of South Africa and published in Vol. 48, No. 11, of the journal of the society. Soluble Salts. In flotation circuits it is seyiom necessary to analyze the solution for soluble salts because higher concentrations of impurities, within the limits of the alkaline circuits employed, can usually be tolerated. In cyanidation work, however, the serious effects of cyanicides have already been discussed. Poor extractions are not usually the direct result of the presence of such salts, but they are indicative of elements in the ore that consume cyanide before the gold can be dissolved and necessitate abnormally large additions of this not inexpensive reagent. Cyanide solutions are therefore frequently tested, especially during starting-up periods for such elements as sulphur, copper, and iron or the thiocyanates, ferrocyanides, etc., determined by direct titration (see Appendix). In addition to the problem of high cyanide consumption, the presence of soluble elements in cyanide solutions that interfere with precipitation can seriously affect metallurgical efficiency and costs. High concentrations of zinc and copper are well known for their inhibiting effect on gold precipitation with zinc, but relatively low concentrations of nickel and thiocyanate, such as are encountered at the Kerr Addison Gold Mines, Ontario, Canada, can also cause precipitation difficulties. The serious effect of chromium is discussed under "Cyanicides." High concentrations of arsenic resulting from the treatment of roasted arsenopyrite concentrates can inhibit precipitation completely. Precipitation. An article appearing in Trans. 50, CIM. and M., 558, 1947, "An Apparatus for Comparing Various Zinc Dusts for Gold and Silver Precipitation" by D. J. A. Dahlgren, describes a method and device that may be useful to mill operators when problems of precipitation arise. Since flows of pregnant solution as low as 4 liters per hr. are handled, the apparatus has possibilities as a precipitation unit for use in cyclic laboratory cyanide tests. Provision is made for de-aeration, the metering of the solution at a constant head past an orifice through which a regulated suspension of zinc dust is fed by means of a novel reciprocating feeder, filtration, and the metering of filtration and control of vacuum. The efficiency of precipitation is obtained by dividing the number of ounces of gold or silver precipi-

234


tated per ton of solution by the number of pounds of zinc dust fed per ton of solution. THE PROBLEM OF BARREN SOLUTIONS

In most operating plants the loss of barren solution in final thickener underflows (straight C.C.D. plants) or as moisture in final-stage filter cakes is sufficient to maintain a solution balance that prevents the accumulation of impurities beyond the critical conce~tration which might interfere with precipitation. In other cases, it is necessary to discard continuously a certain proportion of barren solution from the system, with or without "flood washing" of filter cakes, which at least ensures very thorough displacement of gold values. The disadvantage of this scheme is that it increases the loss of residual free cyanide and lime over and above that normally lost through filter cake or thickener-underflow moisture content. To give an idea of typical reagent consumption and the compositions of barren solution, refer to Table 35, which was prepared from data received by the Dorr Company in 1941 in response to a questionnaire on the subject that was submitted to a number of operating companies. AUTOMATIC RECORDING AND CONTROLS

In recent years much attention has been given by certain progressive operators to automatic recording of the various critical factors in mill control and even to fully automatic regulation of tonnage, reagent, and water flows, in the interest of smooth performance, higher operating efficiency, and the elimination of tedious routine testing and manual control. .. The November, 1947, issue of E. and JJ1.J., pp. 136 et seq., describes a number of the more modern devices. The following account is taken in part from this issue and in part from Taggart's "Handbook of Ore Dressing" and other publications. Crushing and Ore Handling. While crushers have for many years been protected by magnets that remove magnetic iron from the ore stream, nonmagnetic manganese steel may pass by to wreck an expensive machine. A new electronic tramp-iron detector (described in E. and JJ1.J., August, 1945) mounted over the belt and suitably connected through relays will stop the drive motor if any steel, magnetic or otherwise, is on the belt. Malcolm Black in "Wright-Hargreaves Improves Its Milling Practice," E. and JJ1.J., March, 1939, describes the use of a photoelectric cell placed on top of and at one side of a surge bin supplying the ball mills and directly opposite a light source. When the ore builds up to a point where the light ray is intercepted, the power is automatically cut off from the motor on

235

PLANT CONTROL

the belt feeding the bin. Also the conveyors that feed th~ ball mills from the surge bin are each provided with a device consisting· of a shoe that rides on the surface of the ore stream and is connected to a switch mounted just above the conveyor. 'When the shoe falls below a certain level owing to a blocked chute or empty surge bin, a horn alarm is sounded. "Remote Control a Feature of an Ore Conditioning," E. and .M.J., April, 1947, describes the Tennessee Coal and Iron Company's big cenTABLE

35.

REAGENT-CONSUMPTION AND BARREN-SOLUTION DA'l'A

Name of property designated ne

.,

.Item A

B

-Tons are per day ............. Cyanide (NaCN) consumed, lb. per ton are ............. Lime consumed, lb. per ton ore ........................ . Precipitation rntioj ........... Lead salts added, lb. per ton. Barren wasted, tons per day ... Cyanide (NaCN) loss in tails, lb. per ton ore .............. Analysis barren solvent: ~ Total cyanide (NaCN) ....... Free cyanide (NaCN) ........ Zinc (Zn) ..................... Copper (Cu) .................. Pot. thiocyanate (KCNS) ... Lime, (CaO) .................. Ferrocyanide. .... . . . . . . . . . . . .

1000

1500

C D E --- - - - 4100

4000

F

---

G H - - ---

I

---

2100

225*

,100

4905

1200

1.0

0.9

5

1.64

1.65

8.4

2.0

0.40

0.55

8.0 4.0 0.37 Yes

6.4 6.0 0.04 No

2.3 0.6

2.5 3.5 § 000

0.3 8.3 § Yes

18.0 2

.... No

1.37 0.63 12+ No

No

2.06 1.01 0.01 No

4.0 3.0 No 60

....

0.14

....

0.058

0.14

0.74

0.15

0.064

0.02

2.88 1.50 0.80 2.36 5.20 0.76 0.24

2.5 0.3 0.06 0.8 2.0 1.5 1.5

0.71 0.65 0.00 0.17 1.0 1.25 0.002

0.70 0.70 0.023 0.046 0.30 0.54 0.12

0.50 0.45 0.072** 0.002** 0.88*' 1.02** Nil.

0.025 0.013 0.008tt Nil. 0.12tt 0.002tt 0.009

0.04 0.03 0.005tt 0.03tt 0.10tt 0.10tt 0.05tt

0.30 2.06 0.58 0.30 0.008tt .... 0.035tt 0.06 0.40 .... 0.'10 0.35" 0.011 ....

II

I

* Concentrates.

t Tons of solution precipitated per ton of are milled.

+

Lb. per day to zinc cone. § Lead acetate to clarifier (quantity not stated). II Very small addition. ~ Lb. per ton solution unless stated. ** Grams per liter. tt Per cent.

tral plant in the Birmingham district, where the operations involved in filling and withdrawing ore from forty 800-ton storage bins for blending are directed from a central control room. A similar central control system is also used at the Tennessee Copper Company's London and Isabella mills, where feeders, screens, conveyors, and crushers are controlled by one man at a central station using signal lights and push-button switches. Automatic feeders, such as the Hardinge constant-weight feeder, are extensively used to regulate tne feeding of materials from bins to drawoff conveyors.

236


Taggart points out that some degree of automatic interlocking is important in any crushing plant if serious spills and damage to equipment due to chokes in any part of the system are to be avoided. Grinding and Classification. The operation of wet-grinding mills are in many instances controlled by watching the ammeters on the drive motors. At the Hollinger mill in Ontario, Canada, large illuminated instruments are placed in full view above the operating floor opposite each mill so that the load the mill is drawing can be seen at a glance. Hollinger also uses a revolution counter on the feed belt, which is periodically calibrated. Merrick weightometers placed on the belts feeding grinding mills are used both to measure the total tonnage milled and to keep a close check on the rate of feed. The Hardinge Co. "electric ear" is a radioelectric device adapted to translate differences in mill sound automatically into changes in feed rate. For details of the electrical circuit see U.S. Patent 2,235,928. Various systems are used for controlling classifier performance. The density of the classifier pool can be regulated by a combination of tonnage control and fixed valve nipples through which the dilution water flows under constant pressure. The familiar pulp balance for reading the specific gravity of pulp is being replaced by the Adams- or Masco-type automatic density controllers, which are actuated by the differential pressure required to bubble air through two open-ended tubes of different lengths submerged in the pulp. By mechanically varying the resistance in an electrical circuit which is connected to a motor-operated valve, the pulp density controls the flow of dilution water. At Utah Copper a roller that rides on the ribbon of classifier sand returned by the rakes adjusts a rheostat that controls the classifier rake speed. The Dorrco hydraulic sizer uses the Minneapolis-Honeywell electric pressuretrol-modutrol system for regulating the spigot discharge valves. Flotation. Referring again to the London mill of Tennessee Copper Company, the E. and ]J1.J. article above mentioned shows a central instrument panel that controls the flotation circuit. This panel contains (1) a recording pH meter that shows the pH in the regrind cirotiit that feeds the copper-flotation section, (2) an automatic recorder for xanthate being fed to the bulk rougher cells, (3) a conductivity cell that tests a continuous "ample of mill pulp water, (4) thymotrols for suppling the direct current to the small motors that run the lime and xanthate feeders, (5) a continuous pulp-density recorder for the grinding circuit, (6) a mill pulp temperature recorder, (7) a time-delay control that is adjustable and keeps the lime feeder from overrunning the pH meter.

PLANT CONTROL

237

In the paper "Roasting and Flotation Practice in the Lake Shore Mines Sulphide Treatment Plant," J. E. 'Villiamson remarks: The close control of pH is one essential feature of good conditioning.. The critical measurement is the spray-tower agitator discharge pH. Shortly after the plant was started up, a Beckman pH meter was installed. This meter, used in conjunction with a Micromax recorder, gives a continuous 24-hr. record of the pH of the pulp entering the first turbo agitator. The pH at this point is checked at hourly intervals by a colorimetric determination. Although many pH indicators were tested, it was found that bromthymol blue was the most reliable. Some of the indicators covering the range pH 6 and 7 were quite useless on the solutions from this pulp.

Tailings. Pumps handling tailings and othe~' relatively large volumes of pulp frequently give a lot of trouble when power failures occur. At the Tennessee Copper Company plants, protective relays have been installed, so that in the event of such failures the feed to the pumps is diverted either outside the mill or to special sumps and heavy spillages are thereby avoided.

CHAPTER XIII

Cya1Zicides and Refractory Ores Cyanidation as applied to ol'dinal'y gold and silvel' ores is a l'elatively simple pmcess. When cyanicides (cyanide-consuming elements) are encouniered in small amoumis in the tl'eatment of such. ores, the various schemes all'eady discussed, such as lise of a lead salt 01' wasting barren. solution, can uSlially be resorted to and sliccessful operation maintained. When, howeuer, the pl'oblem concerns the treaimen: of an ore that does not respond to these simple expedients, certain. modifications of the cyanide process must often be considered. In the pl'esent chapter the Va1'i01iS cyanicides and methode of contl'olling them are discussed, as well as the causes of rejraciorines» in certain. ores, ioqeiher with recommended treatment. proceduree. TiVJwl'e these fail, it may be desirable to incorporoie cyanide reqenerctioti 01' roasting into the treatment scheme. IRON SULPHIDES

While the oxidized iron minerals ordinarily have little effect in cyanida" tion, the sulphides-pyrite, marcasite and pyrrhotite-tend to decompose in cyanide solution. Pyrite is the most stable and least troublesome of the sulphides. Flotation concentrates high in pyrite content are frequently cyanided without undue consumption of reagents. Marcasite decomposes more readily than pyrite, and for this reason what follows for pyrrhotite is to a lesser extent true for certain occurrences of marcasite also. Where pyrrhotite is present in an ore, trouble is usually experienced both in regard to cyanide consumption and gold extraction, for the reactionsinvolved tend to reduce both the free cyanide and the oxygen content of the solutions. Pyrrhotite,' which has the general formula Fe mS m + l' differs in composition from iron pyrite and many other sulphide materials, inasmuch as one sulphur atom appears to be loosely held in chemical composition and is easily capable of forming additive compounds, such as sodium thiocyanate, NaSeN, from cyanide. (On "weathering," the mineral readily yields elemental sulphur.) The [FeS] remaining is particularly prone to oxidation, forming ferrous and ferric sulphate, which interact with cyanide to form complex cyanides. These reactions show that pyrrhotite not only is a power1 Jack H. French and Harold Jones, "Reduction Works Practice at Morro Velho, Brazil."

238

CYANICIDES AND REFRAC'l'ORY ORES

239

ful cyanicide but also tends to rob the cyanide solution of much of the oxygen necessary for gold dissolution. . Ores containing pyrrhotite are always difficult to treat satisfactorily by cyanide owing to the easily decomposable nature of that mineral. Pyrrhotite, if. kept dry, is stable but in most atmospheres breaks down rapidly, and in contact with water the rate of decomposition is still more accelerated. The action is essentiully one of parf oxidation, and the products are not dissimilar from those which occur when pyrite or marcasite are weathered. . .. The main difference is that the rate of decomposition of pyrrhotite is markedly greater than that of the other common pyritic minerals ... and larger quantities of ferrous compounds are formed and have to be dealt with thnnis usually the case with nonpyrrhotitic ores.

To overcome this effect various methods of oJpre-aerating such pulps before cyanidation with special attention to alkalinity control have been used in operating plants, with a considerable degree of success. Formation of Thiocyanate. In his discussion of the paper "Cyaniding at N oranda" referred to below, Norman Hedley says: The reason for the lower cyanide consumption obtained with intense aeration is probably connected with the reactions which take place when alkaline sulphide decomposes in cyanide solution. Alkaline sulphide is one of the initial products of the reaction between pyrrhotite and alkaline cyanide solution. This may be detected during the early stages of contact. The alkaline sulphide, in the presence of oxygen, may decompose in two ways simultaneously. On the one hand, by a series of reactions during which various oxidized sulphur compounds, such as thiosulpbate, thioriate, sulphite, and sulphate are formed, e.q.,

+

+ H 20 = + 202 =

2Na2S 20 2 Na2S203 2NaOH

+

Na2S203 2Na2S04

+ 2NaOH + H 20

and on the other hand, to the thiocyanate

2Na2S

+ 2NaCN + 2H 20 + O2 =

2NaCNS

+ 4NaOH

It is suggested that the relative proportions of alkaline sulphide decomposed by the two series of reactions depend on the intensity of aeration.

Use of Lead Salts. For retarding the above decomposition of sulphides in alkaline cyanide solution the addition of a lead compound (the nitrate or oxide, for instance) is frequently found to be effective. Two possible reactions are apparently involved: (1) Any soluble sulphides formed by the rapid solution of the finer particles of iron sulphide are preferentially precipitated as the highly insoluble lead sulphide, and further consumption of the cyanide avoided. (2) A surface reaction takes place on the larger sulphide particles whereby a film of the same insoluble lead sulphide is formed, and this protective coating inhibits further reaction. Formation of Ferrocyanides. Where the pH of a pulp containing pyrrhotitic mineral is allowed to drop and oxidation is permitted to pro-

240


ceed without sufficient protective alkalinity, the FeS is converted to FeS04 which reacts with the free cyanide present to form ferrocyanide: FeS04 Fe(CN)2

+ 2N aCN + 4NaCN

Fe(CN)2 + Na 2S04 Na 4Fe(CN)6

"Cyaniding at Noranda." An important paper with this title was published in 1946 by C.· G. McLachlan, H. L. Ames, and R. J. Morton (Trans. 49, C.I.JliI. and Jl1.). It describes the successful working out of a method for cyaniding a pyrite concentrate produced by floating the tailing from the copper-flotation circuit. The sulphide content of this feed was 99 pel' cent, of which 10 to 20 per cent was pyrrhotite and 0.3 to 0.7 per cent chalcopyrite. There was no free gold. The exposed gold was attached to pyrite grains or in fine veinlets, some particles less than 1 micron in diameter. There was available for dissolution about 0.05 oz. per ton of gold. In the course of several years of laboratory and plant testing the following conclusions were reached as to the best operating conditions: 1. The economical limit of fine grinding was 90 per cent minus 325 mesh. 2. A well-aerated pulp saturated with calcium sulphate, using a minimum alkalinity of 1.0 lb. per ton of calcium equivalent.. 3. Cyanidation for a period of 6 to 8 hr. at 50 pel' cent solids with a cyanide strength of not more than 0.35 lb. free N aCN per ton of solution.

The problem had two aspects: a study of conditions in the flotation circuit ahead of cyanidation that would provide maximum elimination of pyrrhotite and chalcopyrite" and of conditions within the cyanide circuit itself that would overcome the effect of.copper and sulphide sulphur. The chemical reactions involved have already been discussed. Sub-Nigel Practice. In the recent paper "Treatment of Gold Ore containing Pyrrhotite at Sub-Nagel, Limited" by King, Clemes, and Cross (Jour. C.M. and M.S.s.A., February, 1947) the authors describe investigations carried out to determine the cause of high residues at the aforesaid property and the measures taken to correct the difficulty. The following treatment was evolved: Pre-aeration in three 33- by 48-ft. Puchuca tanks in series, followed by cyanide treatment in eight 50 ft. diameter by 12 ft. side by 4 ft. cone mechanical air-lift agitators arranged in two rows of four tanks in series. The period.of pre-aeration is approximately 12 hr., and cyanide treatment 42 hr.; in the winter months, however, it is usually possible to reduce the time of treatment by taking one to two cyanidetreatment agitators out of the circuit. With the exception of the cyanide present in 2 The flow sheet developed at N oranda is an interesting example of the use of differential flotation to eliminate cyanicides. Had this not been done, cyanidation would not have been practical.

CYANICIDES AND REFRACTORY ORES

241

the solution used for transfer and dilution, the first addition of cyanide is made after pre-aeration, when the strength is brought up to approximately 0.02 per cent N aCN. It has been found essential to keep the alkalinity of the pulp low during pre-aeration and cyanide treatment, and to this end alkalinity in the thickeners is maintained at about 0.002 per cent CaO in the summer months and 0.004 to 0.005 pei· cent in the winter, depending upon conditions of settlement in the thickeners. After the addition of precipitated and by-passed unprecipituted solution to bring the specific gravity of the pulp to 1,40 (45 per cent solids), the alkalinity leaving the pre-aerators is about 0.002 per cent CaO, equivalent, with normal buffer action of dissolved zinc, etc., to IiI-I 9.6. No further lime is added to the pre-aeration and treatment circuit until the last agitators, when the alkalinity is brought up to about 0.01 per cent CaO for the sake of filtration and precipi tation. Alkalinity in the thickeners and treatment circuits is determined by the ordinary titration method, but the main control is obtained by frequent checking with a glass electrode pH meter. A constant-reading pH meter in the pre-aeration agitators would be preferable, but there appear to be technical difficulties in designing a reliable meter to work in agitated cyanide pulp. Milk of lime and dissolved cyanide are fed by independent pumps to the various points in the circuit where additions are required, the quantities of solid lime and cyanide to the mixers being controlled by electric vibrating feeders. The beneficial effect of pre-aeration and treatment in solution of low alkalinity is considered to be twofold; there is possibly complete oxidation of a relatively small portion of finely ground pyrrhotite which has already come nearly to that state from exposure to the atmosphere, particularly during grinding, and the remainder is inhibited by a coating of the products of oxidation formed during the period of preaeration. The hypothesis of inhibition appears the more tenable of the two, since analyses of residues have shown that there is little or no reduction in the pyrrhotite content of the ore during cyanide treatment. In either case, however, the intensive pre-aeration ensures that an excess of oxygen is supplied, so that sufficient remains for efficient cyanide extraction. The residues were reduced from an average value of 0,473 dwt. per ton for a 6 months' period during 1939 and 1940 before the improved scheme was used to 0.278 dwt. per ton for a similar period in 1941 and 1942 after installation of same. Cyanide consumption also fell from 0.93 lb. N aCN per ton to 0.78 lb. per ton over the same interval.

Salsigne Process. The Mines et Usine de Salsigne in the south of France is cyaniding 100 tons a day of a complex gold ore containing besides arsenopyrite, pyrite, and pyrrhotite, sulphides of bismuth, lead, zinc, and nickel as well as small amounts of chalcopyrite. A method of prelime treatment was worked out by M. H. Carron, professor of metallurgy at Delft University, and J. D. Grothe of the Dorr Company. This consists of grinding in water, adding a predetermined amount of lime such that after 24 hr. agitation in Pachuca tanks the solution has dropped to an alkalinity of about 0.01 per cent CaO, filtering, washing, and following this pretreatment step with conventional cyanide treatment. This scheme of treatment made it possible to reduce the cyanide consumption to less than one-half the figure obtained by direct cyanidation.

242


Grothe ascribes the beneficial effects of this prelime treatment to three factors: (1) the removal of readily soluble sulphur in solution, (2) the oxidation and slow precipitation of ferrous iron compounds, "which is probably responsible for (3) the rendering of the remaining sulphides passive to attack by cyanide. It is noted, for instance, that the iron content of the solutions increases rapidly as soon as the alkalinities get very low. Thus the conditions which favor a rapid precipitation of flocculant Fe(OHh do not prevail; on the contrary a slow precipitation can be expected as the pH drops to the range of 8 to 9, and it is presumed that this results in the formation of a colloidal film of iron hydrate which would adhere tenaciously to the sulphide surfaces and account for the passivity effect. We have all observed the formation of such colloidal films on the walls of glass containers. Hornestake Treatment. Among the minerals containing iron in the ferrous state are pyrrhotite, chlorite, cummingtonite (an iron-magnesium amphibole), and a dolomitic carbonate. Some of these carry as much as 30 per cent iron. Arsenopyrite and pyrite are also constituents. The pyrrhotite oxidizes rapidly, perhaps it would be more precise to say steadily, as the action continues seemingly throughout the treatment cycle. In oxidizing it acts as a cyanicide, producing thiocyanate, and withdraws oxygen from the working solutions. Cyanidation should be applied promptly after comminution, and oxygen must be supplied to maintain dissolution of the gold. The gold occurs both in the free state and associated with the sulphides, so that amalgamation is used, followed by cyanidation after grinding to finer than 80 mesh to release the values. The sulphide minerals yield their gold readily, but special attention has to be paid to the chemical and mechanical preparation of the pulp for cyanidation. Up to the present time crushing in water, rather than in cyanide solution, has been practiced, followed by addition of lime and subsequent aeration prior to cyanidation. Tests indicate that aeration must be carried to the extent of low alkalinity; presence of excess lime or addition of more lime after aeration interfere with gold dissolution. Direct cyanidation by grinding in solution is unsatisfactory, with poor recovery and high cyanide consumption, unless addition of lead compounds such as litharge is practiced. With such additions results are more favorable. Much less thiocyanate is formed, and cyanide consumption is markedly reduced, while with identical lime feed the alkalinity of the working solutions is increased. This leads to the belief that some sort of film must be produced on the pyrrhotite which acts as an oxidation inhibitor. The addition of very small quantities of mercury compounds also appears to be of some benefit in increasing the rate of dissolution of the gold.


243

This is in accord with the past experience at Liberty Bell in the years 1913-1919, as reported by Charles A. Chase ( E. and JJ!I.J., Vol. 129, No.2, Jan. 23, 1930). Nathaniel Herz, chief metallurgist, points out that the principal factors in cyanide consumption and the extraction of gold from the Homestake ore include cyanide concentration, oxygen content of the pulp, hydrogenion concentration (pH), and temperature. While the actual oxygen content of the pulp is difficult to determine, since filtration affects the result, ordinary aeration seems to provide the necessary oxygen demand. It was found, however, that the usual titration of alkalinity was unsatisfactory and pH determination was necessary for this purpose. By controlling cyanide strength, pH, and temperature, the results are systematic and readily reproduced and can be interpreted scientifically. The control of the temperature of cyanide solution is the subject of U.S. Patent 2,220,212, Clark, Herz, and Adams. It is shown that the reaction between cyanide and the sulphides present in an ore is accelerated by increase in temperature. While the rate of dissolution of gold is also increased, this effect is more than offset by the oxidation of the sulphur compounds and utilization of available oxygen. It is proposed to cool the working cyanide solution to 35 to 50°F. for optimum results consistent with the cost of plant and' operation involved. Morro Velho Process. "Control of Alkalinity of Cyanide Pulps" by T. Haden, Institution of Mining and Metallurgy, London, Feb. 19, 1941, contains a very interesting account of a difficult cyanicide problem which was solved by the use of buffer salts and careful pH control. It was found that at alkalinelevels above pH 10, the cyanide consumption was reduced but the gold extraction was poor. The formation of thiocyanates indicated an attack of alkali on the pyrrhotite. Below pH 7.0 the destruction of cyanide was very great, with the formation of ferrocyanides and HCN gas. The best gold extractions and lowest cyanide consumption were reached when no ferro cyanide was found in the solution after 6 hr. In the presence of the optimum lead concentration, a starting pH of about 9.6 using about 2 lb. lime per ton, in the presence of suitable buffers, gave the best results. Soluble lead and mercury salts alone were first tried as buffers, but later it was found that zinc salts returning with the barren solution from the extractor boxes (60 per cent return) could replace about onethird the lead demand. Maximum aeration was found essential to high gold recoveries, but this, of course, increased lime consumption. The conclusion was reached that in Morro Velho ores the presence of free hydrate was not essential but properly buffered solutions were.

244


Starting, for instance, with a low alkalinity (trace to phenolphthalein) the solution too quickly fell below pH 7.0. Soda ash, borax, trisodium phosphate, and sodium and calcium acetates were all tried, the phosphate giving perhaps the best result. Lightly buffered, an alkalinity of pH 9.6 at the start of the agitation period would fall to pH 7.6 to 8 in the course of treatment. In unbuffered solutions, the hydrates, subject to rapid alteration, are quite unstable. COPPER MINERALS

Copper is probably the most active and one of the most troublesome of the cyanicides. T.P.494, U.S.B. of 111., gives the results of a series of tests to determine the solubility of copper minerals in cyanide solution. They show that asurite, chalcocite, cuprite, malachite, and finely divided metallic copper are readily and completely soluble under the ordinary conditions of cyanidation. Bornite is largely soluble under the same conditions, over 90 per cent of it being dissolved in warm cyanide solution in 24 hr. Enargite and tetrahedrite are soluble enough to cause excessive cyanide loss and fouling of solutions with arsenic and antimony. While chalcopyrite is not normally very soluble, the experience at N oranda showed it to be an active cyanicide when finely ground. In recent years it has come to be realized, however, that the copper cyanide complexes do have a considerable solvent action on gold. The paper by Norman Hedley and D. M. Kentro, "Copper Cyanogen Complexes in Cyanidation" (Trans. 48, C.I.M. and M., 237-251, 1945) describes investigations carried out by the Ore Dressing Laboratory of the American Cyanamid Company into the loss of dissolving power of cyanide solutions carrying copper. It is generally accepted that copper dissolved in cyanide solution exists in the form of complex ions such as Cu(CN);, Cu(CN);-, and Cu(CN)~--. It is thought that the first of these is the most common form, although at least one authority considers that CU2(CNS)2 copper thiocyanate is the complex most frequently encountered in working solution. While most operators agree that the dissolving power of copper-bearing solutions is not seriously interfered with if the free cyanide strength is kept high enough, the conclusions reached in the present investigation, which was carried out on copper-bearing ores using cyclic tests with mill solutions and later confirmed in actual practice, are 1. The ordinary silver nitrate-KI method of titration is not a measure of the effective free cyanide present. 2. The total cyanide, as determined by the distillation method (see Appendix B), rather than the free cyanide present is a more correct measme of the effective dissolving power of the solution.


245

3. A ratio of total cyanide to total copper of at least 4:1 must be maintained if serious loss of dissolving power is to be avoided. As pointed out by the authors, this is a practical method of operation only in so far as maintaining such a ratio is economic. If -tho cost in cyanide consumption becomes excessive, then the solution must either- be discarded or regenerated. At N oranda a lower ratio, agreeing more nearly with the cyanide-tocopper ratio of 2.3: 1, in the complex N a 2Cu(CN) 3 was found to be effective. In his discussion of this discrepancy, Hedley points out that, since the solvent action on gold is evidently due to the slight concentration of CN resulting from the dissociation of the sodium cuprocyanide, the presence of any other compounds in the ore which disturbed that equilibrium would have a marked effect on the total reserve of CN ions available for gold extraction. Lake Shore. The principal cyanides in the calcine are copper, ferrous iron, and smaller amounts of Mo, Co, Ni, Mn, etc. Most of the cyanicides can be removed by water washing, but this step cannot be used owing to the loss of about $0.70 in gold resulting from the use of salt in the roast. The cyanide consumption during continuous treatment was about 12 lb. per ton of calcine; after batch treatment was introduced, it dropped to about 9 lb. pel' ton, and now by the use of "superaeration," which apparently converts part of the copper content to a form less soluble in cyanide solution, the consumption has been reduced to around 6 lb. per ton. ZINC MINERALS

Referring again to T.P. 494, U.S.B. of M., the following summary of experimental data on zinc in cyanidation is of interest: 1. The zinc minerals smithsonite, hydrozincite, zincite, and calamine are soluble enough under the usual conditions of cyanidation to cause rapid accumulation of zinc in the solution unless special precautions are taken to remove the dissolved zinc. Willemite, sphalerite, and franklinite are dissolved more slowly. Commercial zinc dust is also readily soluble; therefore an excess over that necessary to precipitate the metals from cyanide solutions should be avoided. 2. When zinc dust or zinc minerals are dissolved in cyanide solution, 1.5 to 4.0 lb. of sodium cyanide is used for each pound of the zinc dissolved. 3. The various zinc cyanide compounds formed in the mill solution are only weak solvents for the precious metals even if the usual titration with silver nitrate shows excess free cyanide; i,e., if considerable zinc is present in the solution, the titration for free cyanide is misleading as to the efficacy of the solution as a solvent for silver and gold. 4. The addition of excess lime or caustic soda improves the extraction with solutions containing zinc by liberating free cyanide from the double salt.

246


5. If most of the cyanide present in a solution is combined with zinc, more free cyanide must be added to this solution than to a fresh solution for the two solutions to have equal activity as solvents. 6. The deleterious effect of zinc in a solution may be entirely overcome by using solutions excessively strong in free cyanide. The continued addition of the required amount of free cyanide would soon result in an excessive amount of cyanide being tied up in the mill solution. Furthermore, strong cyanide solutions consume more zinc during the precipitation of precious metals, thereby increasing zinc consumption and causing additional fouling of the solution. It is therefore advisable when the mill solution becomes foul with zinc either to remove the zinc and regenerate the cyanide or discard the solution.

The solution of zinc probably takes place according to the following equation:

According to Hamilton, if sufficient free alkali is present, the double salt is decomposed, yielding alkali cyanide and an insoluble alkaline zincate.

A second reaction, by which the double cyanide is broken up with the generation of free cyanide, is caused by soluble sulphides formed during contact of the solution with the ore:

These reactions and elimination of soluble zinc with mechanical losses of solution in filter cakes, etc., normally prevent zinc from building up to serious proportions in the circuit. If, however, zinc is used to precipitate solutions high III copper (see cyanide regeneration), the accumulation of zinc becomes so great as to cause rapid fouling of solution. LEAD MINERALS

Galena (PbS), Anglesite (PbS04) .3 The two lead minerals behave very similarly toward cyanide, as the reactive effect of galena is largely due to the ease with which it oxidizes to sulphate. It is important to use low concentrations of alkali or lime; otherwise excessive amounts of alkaline plumbite will be formed, which will interact with the cyanide present to form very basic insoluble lead cyanide. 3 R. J. Lemmon, "Reaction of Minerals in the Cyanidation of Gold Ores," D.E. and M. Rev., March, 1940.


For example, 4NaOH

+ PbS04 =

Na2PbO z + Na 2S04

+ 2H

20

and With low concentrations of alkaline plumbite, a less basic lead cyanide is formed. The less basic lead cyanide hydrolyzes with liberation of HCN, which reforms alkaline cyanide with the excess alkali present. Thus, 2Na2Pb02

+ 2NaCN + 3H

20

= Pb(CNh PbO

+ 6NaOH

and It has been observed! that, if free gold is roasted in the presence of sulfur and lead salts, or if pyritic gold ores containing even a small percentage of lead minerals are roasted to free the gold from the pyrite, lead compounds coat the gold particles enough to make them almost insoluble in the usual mill cyanide solution containing lime for protective alkalinity. Also, the lead compounds remaining in the calcine are slightly soluble in the mill solutions, and such dissolved lead may retard the dissolution of the gold by the removal of oxygen or prevent contact between the gold and the solution by precipitation in some form on the gold surface. Direct cyanide treatment without any protective alkalinity will dissolve this coated gold in time, but the cyanide loss in the treatment of such refractory ores may be prohibitive.. It is shown that high gold recovery can be obtained from such calcines by preliminary treatment with acid brine to remove a large percentage of the lead, followed by cyanidation for the recovery of the gold. There is the possibility that by close control of caustic used for protective alkalinity in the cyanide solution, so as to keep the pH of the solution under 11, it may be practical to leach such calcine by direct cyanidation without excessive cyanide loss. Satisfactory results have been obtained from the application of these methods on a laboratory scale to two important commercial ores that have been refractory to all other known leaching methods. No more recent references are to be found on this subject.

CHROMIUM MINERALS

Lead Chromite (PbCr04).5 The lead chromate mineral crocoisite is destructive to cyanide in the presence of high concentrations of caustic ,1 "Cyanidation of Calcined Gold Ores Made Refractory by the Presence of Lead Minerals," A.I.Jll.E. T.P., February, 1933. 6 Lemmon, op. cit.

248


alkalis such as lime, by reason of the highly oxidizing salts formed, leading to the production of cyanates, formates, etc., the reaction proceeding thus: 2PbCr04

+ Ca(OH)~

=

PbCr04, PbO

+ CaCr04 + H~O

Further reaction may be partially expressed thus, CaCr04

+

NaCN

+

2H~O

= NaCNO

+

Ca(OH)~

+

Cr(OH)~

It is therefore indicated that free alkalinity in the cyanide solution should be at a minimum. In a paper "Chromium in Cyanide Solutions," Jour. C.111. and 111.8.8..1.., June, 1936, H. D. Bell cle~cribes an investigation into the cause of serious precipitation trouble encountered on one of the Transvaal gold-mining properties. The results of this work were summarized as follows: (1) The metallurgical difficulties experienced were due to the presence of chromium in the cyanide solutions in the form of chromate. (2) The yellow precipitate (which was obtained by the addition of lead nitrate to the plant solutions) was lead chromate, carbonate, and hydrate. It did not contain gold, but its presence in the solutions prevented zinc-dust precipitation of gold from solution. (3) The source of the chromium was the particular f'eldspathic and pyroxine ore. (4) All of .the chromium must be precipitated by the addition of a solution of a lead salt and the resulting precipitate removed prior to de-aeration when effective precipitation with zinc is obtainable. (5) Addition of a soluble lead salt to the cyanide solution gives rise to the formation of hydrocyanic acid, which is readily "fixed" on contact with free alkali.

REFRACTORY ORES

Where gold is associated with arsenic, antimony, tellurium, and certain other minerals, it is often only partly soluble in cyanide solution. The difficulty is usually more of a physical than a chemical one but special methods of treatment may be required. ARSENIC ORES

Mispickel (Arsenopyrite: FeAsS).6 When this mineral occurs 'in a gold ore, a proportion of the gold is frequently in intimate association and only rendered open to cyanide attack by previous roasting. With a "sweet" roast, the calcine usually yields a high gold extraction with low lime and cyanide consumption. Mispickel in the raw condition in an ore, when agitated with strong concentrations of lime or other alkali in the presence of air, can yield alkaline arsenites, thus (FeAsS)~

+ 2Ca(OH) + 130

= 2FeS04

+ 2HCaAs0 + H~O 3

This reaction is a strong deoxidizer, inimical to gold solution by cyanide; 6

Ibid.


249

therefore minimum amounts of lime and cyanide should be employed in order to delay such reaction provided the gold is substantially free. If high concentrations of lime are used, the associated gold may be released, but difficulty may ensue in dissolving the gold owing to lack of oxygen. AN'rIMONY ORES

Stibnite (Antimony Sulphide, Sb2S.). This mineral in an ore acts as a weak acid, combining with alkalis to form salts, of the order of MHSbS. and M;,'sSbS 4 , alkaline sulphantimonite and sulphantimonate, especially in the presence of iron sulphides, using high concentrations of lime or alkalis. It is therefore indicated that minor amounts of lime be used in solution, sufficient to provide a permanent alkaline pH for protection of the small amounts of sulphantimony compounds formed, otherwise attacked by the latent acidity of the ore to liberate antimony pentasulphide and hydrogen sulphide, which would act as powerful cyanicides. Use of low-cyanide-strength solutions is also advisable in order to minimize the speed of disequilibrium of the alkaline cyanide ions. The ore should be pulverized only to a sufficient degree to render the gold particles open to cyanide attack, thus presenting a minimum stibnite surface area for interferent reactions. Some stibnite gold ores are more amenable to eyanidation when previously subjected to surface oxidation in the pulverulent condition, with the formation of a sulphate film, which on contact with water yields an insoluble basic sulphate. The oxidation may also open up the fissure planes, setting free the associated gold. This oxidation may be secured by "weathering" the ore pulp, but in some cases, the agitation of the wet pulp with small amounts of lime and sodium peroxide, together with injected air prior to the addition of cyanide, may be effective. If an antimonial ore or concentrate is roasted before cyanidation, this calls for meticulous temperature and air control, as reduction to the molten metallic condition can occur, with enfoldment of the gold particles. Hillgrove, New South Wales," The gold ores from this section are of special interest due to the gold's being in association with stibnite and in some places with scheelite. When ordinary water-slaked lime was used in the cyanide treatment of tailings from previous milling operation, the gold extraction was fair but the cyanide consumption was 12 to 14 lb. per ton, but by the use of ordinary air-slaked, or agricultural, lime the consumption was reduced to 3.5 lb. per ton. The present lOG-ton treatment plant consists of agitation and decantation equipment-an 18-ft.-diameter mixer vat with paddles and four agitators. Eight-ton 7 Mine and M~ill World Digest, April, 1938, condensed from E. and M. Rev., Melbourne, Australia.

250


batches are fed to the mixer. Sufficient lime is added for the propel' degree of alkalin ity. Then additional cyanide solution is pumped in to bring the strength to 0.03 pel' cent. The batch is then stirred for 4 hr. and then discharged into an agitator vat. The mixer handles two batches pel' day. The four agitator vats are 16 ft. in diameter by 8 ft. deep, fitted with stirring gear. Before running a batch into an agitator, a large quantity of sump solution is first pumped in, and the proportion of solution to solid is kept high. After receiving two batches from the mixer and given a double wash, the agitators are discharged, hosed clean, ready for the next batch. The clear solution drawn off from an agitator is run to two settling vats 16 ft. 6 in. in diameter by 4 ft. deep. The solution stands in settlers until any cloudiness has settled. It is then run through a sand filter and then to the zinc boxes. An 80 pel' cent gold recovery was made by the method of treatment described.

It is generally known that antimony can be readily volatilized, and by roasting stibnite (Sb2S 3) in an oxidizing atmosphere, antimony oxide can be driven off and condensed. At the same time, anyone who has had experience with such roasting knows that it is a delicate process. To begin with it is impossible to drive off the antimony completely. Furthermore, unless temperature control is very exact, the ore being roasted is liable to be fused into a solid mass, entirely hopeless as regards subsequent roasting or treatment." Provided the roasting is carefully done, the scoria remaining, which contains the gold, also contains a residual amount of antimony which still stands in the way of normal cyaniding, as the gold is, if anything, still more intimately associated with it. The treatment of the calcines with an alkaline leach for the conversion of the residual antimony oxides to soluble antimonates has been proposed as a method for overcoming the difficulties usually encountered in attempting to cyanide such material. At the plant of the Consolidated Murchison Co. in the Transvaal, South Africa, where a refractory high-antimony gold ore is being treated, flotation and cyanidation of the flotation concentrates and tailings are being practiced. The association of the antimony and gold is very intimate, and during 1946, while the antimony recovery amounted to 70.6 per cent, only 62.7 per cent of the contained gold was extracted. TELLURIDE ORES

Solubility of Tellurides. In Bul. CIM. and M., June, 1933, under the heading "Gold Tellurides Are Soluble in Cyanide," W. E. Johnston wrote: The gold-bearing tellurides behave very much as does gold itself. Finely divided particles are dissolved fairly rapidly; coarser particles more slowly. All tellurides B There is reason to believe that use of the FluoSolids roasting technique with its special features of close control of temperatures and gas composition may improve the metallurgical treatment of antimonial ores.


251

are soft and brittle and also have high specific gravity, with the result that they are extremely finely divided in ordinary mill practice. The ordinary high-lime solution of Kirkland Lake practice is suitable for dissolving of the gold tellurides. Oxygen is necessary but apparently not more so than in dissolved gold. In E. and M.J., August, 1933, Johnston returns to the problem with the results of experiments. Before detailing these he says that a search of the literature failed to show that anyone had actually conducted cyanidation tests on either specimen tellurides or those concentrated from an ore-meaning Kalgoorlie, Cripple Creek, and Kirkland Lake. Experiments with cyanide on these ores gave the following conclusions, with which A. L. Bloomfield, H. V. Wallace, and John Dixon of Kirkland Lake substantially agree: 1. Gold-bearing tellurides do yield their gold to cyanide if they are in a finely divided state and excess lime is used. oJ 2. Sodium peroxide greatly reduces the time of treatment required for maximum extraction. It is not beneficial when used in quantities equivalent to commercial use. 3. The tellurides are very brittle and, owing to their high gravity, will be retained in the mill circuit for a long time. They will thus be in a finely disseminated state, approximately minus 1600 mesh or the size required to yield a maximum extraction of their gold. 4. Up to the present, gold-bearing tellurides have not been found in large quantities in mill tailings or in concentrates recoverable from them.

As in the case of many arsenopyrite ores, however, those carrying tellurium and sulphotellurides often require roasting in order to open up the gold content to attack by cyanide solutions. GRAPHITIC ORES

The so-called "graphite" or carbonaceous mineral that exists in certain ores in California, Montana, West Africa, and western Australia has been a source of considerable difficulty in cyanide plants for many years. It usually occurs in graphitic schists adjacent to the ore bodies and with quite an irregular amount of carbon, increasing the difficulty in plant control. Several different treatment methods have been developed to overcome the premature precipitation of gold by the carbon from cyanide solutions. In West Africa, a posttreatment was given following usual cyanide practice by releaching with sodium sulphide solution, which acted as a solvent for the gold precipitated by the carbon. Results, however, were not satisfactory. The Silver-Dorfman process, in which the ore, crushed in water, was treated with small quantities of fuel oil, kerosene, or a combination of both prior to cyanidation, was successfully used at some plants. The oiled graphite lost, to a large extent, its precipitating power. Roasting was a satisfactory solution, but its cost precluded its use on many low-grade ores.

252


The later development of flotation for gold ores, has, in most cases, provided a satisfactory pretreatment and has been applied in several ways. Flotation removes most of the freed carbon with the concentrate, which is then roasted prior to cyanidation or shipped to a smelter. McIntyre, using flotation with cyanidation of concentrate, refloated their cyanidation tailing and returned this concentrate to the head of the cyanide circuit when they encountered carbonaceous ore. Graphite is present in the gold-silver ore in the Timmins Ochali Mining Co. mill in Columbia." It was noted that: Considerable carbonaceous schist occurs in the run-of-mine ore, not all of which is detrimental. However, small quantities of graphitic schist act as a precipitating agent and are found to reprecipitate values from the solution in the agitating circuit. To remedy this, diesel fuel oil is added at 125 drops pel' minute to the classifier sand discharge. The oil acts as a collector for the graphite, coating and inhibiting its precipitating properties. A collector ring placed around the thickener feed well collects the graphitic froth which floats to the surface. This is skimmed off several times a week and stored. In two and a half years, less than a ton of it has been collected. The small amount of graphite that gets by passes harmlessly through the circuit and goes out to tails.

E. B. Leaver and J. A. W 001£ of the U.S. Bureau of Mines described a number of experiments on California carbonaceous ores in T.P. 481, 1924, and some of their further work on the same subject was published in R.I. 2998, U.s.B. of M., 1930. 9 Wilson and Darnell, "A Lode Gold Mine in Columbia," E. and M.J., Vol. 143, No.5, May, 1942.

CHAPTER XIV

Cyanide Regeneration and Miscellaneous Processes Cyanide regeneration o.ffers a practical means of overcoming the otherwise heavy cyanide consumption freqllently encousuered in the treatment of gold and, especially, silver ores, where cyanicides (cypnide cotisuminq minerals) are present. lVliscellaneolls processes used. for the extraction of gold and silver values by hydrometalhtrgical means are also discussed in this chapter. These inchide the carbon-cyanidation, the bromocqomuie, the ammonia-cyanide, and chlorination procesees. Thouqh. some of these are not used commercially today, they are of interest both historically and because they coruribuie to the Sllm of Ollr general technical knowledge, out of which new and improved processes fOT the flltllTe may be developed.

A. J. Clark, metallurgist of the Homestake Mining Co. for 30 years, once made the significant remark that every process, no matter to what extent it might be regarded as impractical, should be reviewed every 5 years with the thought in mind that improvements in equipment' and techniques might possibly justify its revival as a working scheme. In keeping with this viewpoint, we have included in the present chapter brief descriptions of certain processes generally regarded as out of date and superseded by more modern and technically efficient methods, for the fact remains that in many cases the cost of treatment by our improved present-day metallurgy is often discouragingly high and in other cases it is not possible to treat the ore at all. If the technology of ore treatment is to make a steady advance, it is imperative that research men and engineers keep an open mind in the matter of reappraising the obsolete processes and also what may sometimes appear to be the new-fangled ideas of overenthusiastic inventors. Both may well contain the germs of ideas that lead to important process developments. CYANIDE RECOVERY OR REGENERATION OF CYANIDE

With regard to regeneration of cyanide, J. E. Clennell said m The Cyanide Handbook: Since the main cause of cyanide consumption is the formation of soluble double cyanide or complex cyanogen compounds of the base metals, and since solutions 253

254


highly charged with such compounds are more or less inefficient as solvents of gold and silver, it has been suggested that the cyanogen in such liquor might be recovered in the form of simple alkali cyanides, by treatment with suitable chemicals. This has been carried out in practice in some cases, but generally speaking the cost of chemicals, power, labor, and other charges required for such treatment outweighs the advantage gained by it.

The latter part of the foregoing statement may be questioned by some, but although the few regeneration plants at work report a saving in cyanide and an improvement in the treatment, methods for the recovery of cyanide have not been generally adopted. Under normal operating conditions the mechanical loss of cyanide discharged with the tailing is an important factor in cyanide consumption. This loss results from imperfect washing, which in turn is the result of limiting the water wash to the amount of water necessary to maintain a balance of plant solution. This clifficulty can be practically eliminated by using solution from which the cyanide has been removed to extend the washing period, according to W. E. Crawford of Fresnillo, Mexico. In Handbook oj Ore Dressing, 1928 ed., by A. L. Taggart, R. C. Canby said: Precipitation of gold and silver from cyanide solution by means of zinc or aluminum results in regeneration of cyanide, probably not in the form of 'alkali cyanide as originally added but in a form in which an equivalent amount of effective cyanide ion is present.

The common method for regeneration of cyanide is by acidulation of the solutions. Allor part of the cyanogen is converted thereby into hydrogen cyanide, which is fixed by an alkali (generally lime) and returned to the cyaniding system. This is the Mills-Crewe process, the principle of which is described by C. W. Lawr in T.P. 208, fi.I.l11.E., 1929, in which he also gives a selected list of 38 references to regeneration: PRINCIPLE OF CYANIDE REGENERATION

The solution, be it a' weak wash or a foul solution, is made acid by bringing it into contact with sulphur dioxide. The acidified solution is then transferred to a closed tank in which air and solution are brought into intimate contact. The air leaving the tank charged with hydrogen cyanide is then passed to another tank in which it is mixed with an alkaline SOlution, the latter absorbing the HON and leaving the accompanying air clean for reuse in removing more HON from the acidified soluti on. The extent to which the acidified solution will become impoverished of its cyanide will depend upon the acidity, the amount of air brought into contact with the solution, and the quantity of residual HON left in the air after the latter has passed the absorbing apparatus. (The system is closed so that the same air is used repeatedly.) Of course the amount of air required will depend on how efficiently it is utilized, but where other conditions are equal, it may be stated that the amount of cyanide

CYANIDE REGENERATION AND MISCELLANEOUS PROCESSES

255

removed from a given volume of solution increases with an increase in air, and, further, an increase of air will do more good or a decrease will cause poorer results than almost any other change that could be made in the plant. (By increase in air is meant an increase in the velocity of that being circulated.) . The impoverished acid solution may be wasted or used as a water wash on the filters, either before or after filtering, depending on whether the solution contriins enough silver or copper to pay for its removal. If the spent solution does not contain any silver, but much copper, it is doubtful whether it could be used as a filter wash before filtering, because the slimy nature of the precipitate would affect adversely the filter leaching rate.

The Mills-Crewe cyanide-recovery process is... applicable only in plants of rather large tonnage and best on solutions resulting from the treatment of silver ores or those containing appreciable amounts of copper. Cyanogen in mill solutions exists chiefly as free alkali cyanide, as zinc and copper double cyanides, and as sulphocyanide and ferrocyanides. From the free cyanides and zinc double cyanide, substantially complete recovery of the cyanogen is easily effected. Part of the cyanogen combined with copper is readily recovered; regeneration of the remainder, with the cyanogen existing as sulphocyanide and ferro cyanides, requires special treatment which well may be justified in large plants or in plants where unusually strong solutions are employed. Cyanide Regeneration at Pachuca, Mexico. At Pachuca treatment of the silver ore, which carries some copper, and gold involves a high cyanide consumption, which is reduced by cyanide regeneration. Mter the usual washes of the filter cake, including a water wash passing to mill make-up, the cake is given a further wash of 20 min., and the values in cyanide and precious metals contained in this solution are removed in the regeneration plant, the cyanide being returned as a gas and absorbed in the mill solution, while the silver and gold and quantities of copper are recovered in the form of a precipitate, which is shipped to a smelter. The plant treats 3800 tons per day of a solution composed of 1000 tons of barren solution and 2800 tons of water wash from the filters. Regeneration consists of (1) acidifying the solution with sulphur dioxide to neutralize lime and convert the cyanides to hydrogen cyanide, (2) vaporizing the hydrogen cyanide from this solution by means of a large volume of air, (3) absorbing the hydrogen cyanide carried by this air in the regular mill solutions, (4) adding zinc dust to precipitate the gold, (5) recovering the gold-silver-copper precipitate by filtration. The sulphur dioxide gas used in the process is made in a rotary sulphur burner. It is brought into contact with the cyanide solutions in an aeidifier. The HCN formed in the acidifier remains in the acid solution until removed in the dispersers. This gas is dissolved in the solution, and al-

256


though it can be removed fairly easily, it is fixed sufficiently in the acidified solution so that, upon passage through a weir box 6 ft. in length open to the atmosphere under normal conditions, no loss of HCN can be detected by silver nitrate titration of the solution as it enters and leaves the weir. At times, an odor of HCN is noticeable, indicating a slight loss. The HCN is removed from the acidified solution by bringing a large volume of rapidly moving air into contact with the acidified solution spread over a large surface. Solution-surface exposure is obtained by grids and spraying devices. At least 15 cu. ft. per min. of air per ton of solution treated in 24 hr. is required. The HCN removed from the acidified solution then is absorbed by an alkaline plant solution in horizontal absorbing towers in which the solution and the air containing the gas are brought into close contact. R. R. Bryan,' general superintendent of mill at Pachuca, writes: The regeneration of the quantity of barren solution was undertaken to keep the plant solutions reduced in certain constituents which, when allowed to build up, caused poor extractions and poor settling. Copper is one of these deleterious constituents, and although unknown, we suspect that there may be others. We endeavor to keep the copper content of the thickener overflow below 80 grams copper per ton of solution. With this bleeding of 1000 tons barren solution per day there has been a great improvement in settling and former periods of unusually low extraction have been avoided. Another change in cyanide regeneration has been the introduction of entirely automatic acid control. This is accomplished by means of a Beckman pI-I meter and a Bristol potentiometer controller which operates the air valve on the sulphur burners. The correct pH for optimum results depends on the constituents of the particular solution. For filter washes it is pH 5.6, but with the addition of barren we use pH 5.1. The difference in pH required for different solutions appears to be due to different amounts of double zinc cyanides which they may contain. The double zinc cyanide requires a higher pH for its complete regeneration than do the simple cyanides. The introduction of automatic pH control has been a major improvement in our cyanide regeneration. The amount of sulphur required has been reduced, the tails are lower, and less trouble from liming up is encountered. Before pH control, the regeneration of an considerable tonnage of plant barren solution was very difficult because of liming up. The use of pH control has greatly reduced this difficulty.

Hudson Bay Mining and Smelting Company. The operation of a cyanide regeneration plant in Canada is described by the mill staff in "Cyanide and Regeneration Plant and Practice at Flin Flon" (Trans. 49, C.I.M. and M., 1946) as follows: From the storage tank following precipitation, barren solution is pumped to the Mills-Crewe regeneration plant for recovery of the cyanide, which is reused during agitation. 1

Personal communication to the authors, May, 1948.

CYANIDE REGENERATION AND MISCELLANEOUS PROCES'SES

257

The elements of a single regeneration unit are shown in Fig. 60. They consist of 1. A dispersing tower. 2. An absorbing duct. 3. A blower rated at 36,000 cu. ft. per min. and 4-in. water gauge. 4. A centrifugal pump for dispersed solution. 5. A centrifugal pump for absorber solution. Acid

Ijne~

ABSORBE~-PLAf'll

FIG. 60. Cyanide regeneration unit at FHn Flon, Manitoba, Canada. Acidified barren solution enters the top of the dispersing tower, where it is distributed across the tower section by three launders. The solution drops through the grid-packed tower countercurrent to an air stream discharged by the blower and into a sealed sump at the tower bottom. From the sump, a centrifugal pump transports the solution to the next unit. The air stream in its passage up the tower sweeps the hydrocyanic acid gas, liberated by the action of the acid, into an air duct leading to the absorber. This is a tunnel about 40 ft. long and 4 by 6 ft. in section. One end joins the air duct from the disperser, and the other end forms the blower inlet. The tunnel floor, over which

258


absorber solution flows, slopes at 7~ in. per ft. to 11 scaled sump. From the sump, 11 centrifugal pump elevates absorber solution to 11 storuge tank feeding the absorbers and agitntors. Six rotor sprays are installed at intervals across the tunnel, about 3 in. above the floor. The rotors are 12 in. diameter by 4 ft. long and are driven at 900 r.p.m. by 3-hp. motors. The adjustment of 11 butterfly weir at each rotor controls the depth of its immersion and the fineness of the mist in the tunnel atmosphere. A frumo of 30 metal louvers located across the tunnel near the blower intake protects the Ian from lime-scale ccating by minimizing the mist entering the blower. The air streum carrying hydrogen cyunide gas enters the absorber ut its junction with the disperser air duct. It is drawn through the mist of absorber solution, a we11k lime slurry. The cyanide gas reacts with the absorber mist and remains in solution l1S calcium cynnide, while the other guses pass on into the blower. The blower discharges into the disperser tower, completing the gas cycle. The entire plant consists of 1. Four dispersing units.. 2. Four absorbing units. 3. An absorber storage tank feeding the absorber units and cyanide agitators. The dispersing units operate in series. Barren solution is acidified with waste sulphuric acid from the electrolytic section of the zinc plant in a lead-lined box ahead of unit 1. Acid addition is controlled at this point to give 11 residual acidity of 1% to 21b. acid per ton of solution at the last disperser. A minimum residual strength of 17~ lb. per ton is required for good disperson. From the last disperser, 'the solution is pumped to the copper sulphate plant for further treatment, The absorber units operate in parallel. From the storage tank, the absorber solution flows by gravity to the high end of each of the four units. It becomes enriched in cyanide during its passage through the absorber and is pumped back to the storage tank by 5-in. pumps. An 18- by 18-ft. wood-stave tank equipped with a 48-in. ship-type impeller is used for absorber solution storage. From this reservoir, the final regeneration plant product is drawn for reuse in the agitators and circulates to each absorber unit. The alkalinity of the solution is maintained at Llb. C110 per ton by addition of plant lime slurry. The draw-off to the agitating section is replaced with water. A portion of the flow to the ugitators is used to dissolve and transport the raw cyanide addition. The regeneration operation for a year is summarised in Tuble 36. TABLE

36.

CYANIDE-REGENERATION DATA

Barren solution to regeneration: Daily tonnage . Regenerable cyanide (Nl1CN) . Lime (CaO) . Tailing from regeneration: Regeneratable cyanide (NaCN) . Residual sulphuric acid (H 2SO.j ) . . . . . • • . • • • • • • • • • • • • • • • • • • • • • • • Sulphuric acid (H 2 S0 4) consumption: Barren solution . Recovery of regenerutable cyanide: Percentage , . Cyanide (Nl1CN) per ton cyanide heads " . Raw cyanide consumption: Cyanide (NaCN) per ton cyanide heads .

2814 tons 2.23 lb. per ton 0.25 lb. per ton 0.18 lb. per ton 1.91 lb. per ton 4.58 lb. per ton 91.5 per cent 1.69 lb.

1.15 lb.


259

Leaver and Woolf Process, u.s. Bureau of Mines. In T.P. 494, "Copper and Zinc in Cyanidation. Sulphide-Acid Precipitation," the authors E. S. Leaver and J. A. Woolf describe a process for the regeneration of a large part of the cyanide combined with copper and zinc.in fouled mill solutions. This process, which as far as we know has not yet been used on commercial scale, involves the addition of a soluble sulphide to the' solution followed by acidification in a closed reactor to pH 5 to 0, whereby the zinc and copper are thrown down as sulphides and up to 80 per cent of the combined cyanide regenerated. After removal of the precipitate by filtration lime is added to the filtrate to "Ji'C" the HCN regenerated, and the solution is ready for reuse. oJ The silver would normally come down with the copper and zinc precipitate, but it may be separately removed as silver sulphide if desired before acidification. Only a part of the gold is precipitated; the rest must be recovered by the usual zinc-dust treatment. The process must, of course, be carried out in a carefully closed system to avoid the hazards of HCN asphyxiation, and it has the disadvantage that sulphocyanates, cyanates, etc., are not broken up. It is stated that the precipitated sulphides are readily filtered and contain no cyanide, but the exact manner of handling and retreatment for extraction of precious metals is not described. This technical paper is, however, of special interest to those concerned with the problem of soluble copper and zinc in cyanide solutions. General Engineering Company Process. In this process the free cyanide and free lime as well as any copper, zinc, or silver compounds are precipitated by the addition of chemically equivalent amounts of zinc sulphate. The precipitate, following filtration, is treated with acid for regeneration and recovery of CN by the usual stripping and absorbtion methods, and the filtrate treated with zinc dust or other suitable precipitant for recovery of the gold. The cyanide' tied up with the copper and silver is lost in the proposed smelting method for recovering the metals, but the process has the advantage that only a small bulk of precipitate is acidified as against the whole volume of solution in other processes, and it would also appear that a considerable part of the acid ZnS04 could be reused after neutralization with lime. Electrolytic and Other Methods. Electrolysis of cyanide solutions results in regeneration of the cyanide combined with the metal complexes, but this method has been largely abandoned owing to the high cost of installation and operation. Other methods proposed for the precipitation of the metals dissolved in cyanide solution usually result is some degree of

260


cyanide regeneration, though few have passed beyond the experimental stage. Australian Practice," At the New Occidental Gold mines N.L. Cobar, New South Wales, Australia, solution containing copper from the direct cyanidation of ore containing pyrrhotite and chalcopyrite (from 0.1 to 0.3 pel' cent copper) is treated by a batch process. When the working solution increases in copper content above 0.03 per cent, the double cyanide of copper, CU2(CN)22NaCN, is precipitated by strongly acidifying the solution with H 2S0 4 . This brings down a bluishblack precipitate of cuprous cyanide which is allowed to settle, the clear solution being decanted off into an agitation vat where lime is added to give a protective alkalinity of approximately 0.1 per cent CaO. By this method profitable regeneration of the free cyanide from the double cyanide is accomplished and the copper content of the plant solutions kept below 0.03 per cent copper. It was found in practice that, when the copper content increased above this figure, dissolution of gold was reduced considerably. The low-grade copper precipitate produced in the process is run to waste from the settling tank, since shipping it to a smelter proved uneconomical. Gold-copper Residue Treatment. H. B. Wright stated in E. and M.J., May 5, 1923, that regeneration of cyanide from cupriferous cyanide solutions is profitable, easy of application, and made possible the treatment of a dump of refractory residue in New South Wales, Australia. This material contained about $4 gold per ton and 0.1 per cent copper in the sand and 0.33 per cent in the slime. While treating more than 10,000 tons of this mixture, half of the cyanide was regenerated and about 1 lb. per ton copper was recovered. To neutralize acidity, 4 lb. lime was added to each ton of sand going to the leaching vats; 2 tons of cyanide solution pel' ton of sand was used during the 14 days' treatment. But as it is essential for regeneration to deprive solutions of alkalinity or free cyanide, the sump solution ~itrating 0.06 per cent N aCN and 0.04 per cent CaO was applied to the 70-ton sand vats, a half ton to each ton of sand, preceding the neutralization by lime. The solution draining away was neutral to phenolphthalein and was pumped to the regenerating part of the plant for subsequent treatment. It was found best, by repeated tests, to apply one case or 224 lb. cyanide all at once to-a vat containing 70 tons sand. The next step was to pump regenerated solution of 0.14 per cent NaCN and 0.10 pel' cent CaO strength over the cyanide placed on top of the sand, which dissolved all of the new cyanide required and raised the strength to 0.40 per cent. This solution 2

C.E. and 111. Rev., Jan. 10, 1947, p. 134.


261

was circulated by pump for 8-hr. periods over two days. Then it drained . to the zinc boxes, having 0.15 to 0.20 per cent free cyanide. Precipitation on zinc was for the richer and stronger solutions only, and 80 per cent precipitation of the gold was considered good work under the conditions of treatment. Calcium sulphate mud was a nuisance in the. boxes. . The gold-bearing sludge was so refractory that it was sold to a smelter. Precipitation of gold from the weaker solutions was accomplished by sulphuric acid. Copper was present in the solution as the double cyanide Cl12(CN)22NaCN and was precipitated in the padd]e agitator mentioned in the next paragraph by means of sulphuric acid. The dried precipitate assayed 60 per cent copper, 70 oz. gold, and double that of silver. Regeneration was done as follows: The solutions neutralized and freed from active cyanide, as already mentioned, were eventually pumped to an elevated agitating tank of 50 tons capacity. Enough sulphuric acid was then added to throw down as cuprous cyanide, Cl12(CN)2, all of the copper present. Each charge of solution required from 7H to 14 lb. per ton. The agitator, from which an unpleasant odor arose, was then stopped, and the cuprous cyanide precipitate allowed to settle for 1~~ hr. The clear liquor, charged with hydrogen cyanide, was decanted into a 50-ton Don agitator. Milk of lime then was added until the solution showed 0.09 to 0.13 per cent free CaO and 0.14 per cent free NaCN. Agitation was effected by pumping. The regenerated solution was then pumped into the head tank and reused in treatment, as described. The Don agitator was cleaned of insoluble matter at periods of 3 to 6 months. The cuprous cyanide precipitate from the paddle agitator was allowed to accumulate during six charges, then was washed out and drained on a filter consisting of wooden slats, coconut matting, and sacking and finally dried and shipped to the smelter. NOTE. Although no ill or serious results ever attended the use of this process at this plant, it should be used with caution because of the extreme danger of cyanide poisoning by hydrocyanic acid gas.

MISCELLANEOUS PROCESSES CARBON CYANIDATION 3

Countercurrent carbon cyanidation with simultaneous dissolution of gold by cyanide and its adsorption by carbon offers several advantages over other carbon cyanidation processes as (1) the rate of dissolving the gold is faster, (2) higher grade gold-bearing carbon is obtained, (3) less carbon per ton of ore treated is required, (4) separate dissolving and adsorbing 3

Private communication from Dr. T. G. Chapman.

262


units are not needed, (5) adsorption of dissolved gold by colloids or graphite is reduced, and (6) capital and operating costs of plant are reduced. Pilot-plant operations' since 1940 which employed this method of cyanidation include one plant which operated without agitation and three plants which employed agitation but used different methods of separating the carbon from the ore pulp. Harquahala Pilot Plant. The Eagle Picher Mining and Smelting Company in 1940-1941 erected and operated a 25-ton capacity pilot plant on Harqua Hala amalgamation tailing near Salome, Ariz. The tailing treated contained considerable colloidal material, and treatment of this material by standard cyanide and flotation methods had been tried without satisfactory results. The tailing was pulped with water, lime, cyanide, and activated carbon to a consistency of 70 per cent solids. The pulp was allowed to stand without agitation for a period of 15 to 24 hr., an~l then the carbon was separated from the pulp with a Denver flotation machine. The extraction of gold ranged from 70 to 75 per cent, and the concentrate contained from 25 to 50 oz. gold per ton. The explanation for the completeness of the reactions involved without the usual agitation of the pulp lies in the continuous migration of gold ions to points of low gold concentration which surround the particles of activated carbon. Getchell Pilot Plant with Flotation. In 1946, the Getchell mine! of Nevada installed a 100-ton capacity pilot plant employing flotation for the recovery of the carbon from the are pulp. Lime, cyanide, and activated carbon were added to the ball mill, and the carbon was separated from the pulp with mechanical sub-A flotation machines. On heads assaying 0.10 oz. gold per ton, the extraction amounted to 85.7 per cent, the tailing assayed 0.014 oz. per ton, and the barren solution contained 0.0005 oz. gold per ton. The grade of concentrate, 2.0 oz. gold per ton, was low owing to the presence of a black mineral in certain parts of the Getchell ore body which floated with the activated carbon. Getchell Pilot Plant with Carbon Containers. In 1947, Getchell mine" installed a second pilot plant of 100-ton capacity to experiment with the elimination of the flotation step in recovering the carbon from the pulp. Originally, coarse activated carbon, minus 10 plus 30 mesh, was used, and the coarse carbon was confined in revolving screens which were partially submerged in the ore pulp. Three agitators were used in series with one 4- by 4-ft. screen, 30-mesh stainless-steel cloth, in each of the first ·1 Hardy, R. A., "Carbon-cyanidation mill tests at the Getchell Mine, Nevada," American Mining Congress, Denver meeting, September, 1946. 6 R. A. Hardy and F. W. McQuiston, "Developments of Carbon by Cyaniclation of Getchell Mine, Nevada," American Mining Congress, El Paso meeting, October, 1947.


263

two agitators and two screens in the third agitator. The. coarse carbon was advanced countercurrent to the direction of pulp flow from agitator 3 to agitator 1. The carbon discharged from agitator 1 to the refinery assayed 82.6 oz. gold per ton. In the latest (1949) modification of this process, five Dorr agitators are used in series and the revolving screens are suspended vertically in the' pulp. In this new scheme the carbon-loaded pulp is on the outside of the screens and is transferred from one agitator to the next by air lifts, while the carbon-free pulp which passes through the revolving screens is passed by air lifts into the preceding agitator countercurrently to the carbon-pulp flow. Thus, pulp which is largely barren of gold leaves the system at one end, while the activated carbon with its load of gold (approximately 400 oz. per ton) is withdrawn from the other end, separated from entrained pulp in a small washing trommel and passed on to "desorption" or step of removal of the gold (see page 265). Barren solutions as low as 0.001 oz. per ton have been reported. In this particular adaptation of the process cyanidation is carried to completion before the carbon is added. American Cyanamid Aerochar Precipitant. In 1948, American Cyanamid Company employing the pilot-plant unit of the Eagle Picher Mining and Smelting Company at Sahuarita, Ariz., conducted continuous tests employing magnetic activated carbon for the adsorption of gold. The carbon was prepared by mixing magnetite with activated carbon and sodium silicate binder. After drying, the magnetic portion was separated and introduced to the ore pulp. After the completion of dissolution and adsorption of the gold, the carbon was separated from the pulp with wet magnetic concentrators. Recovery of Adsorbed Gold and Silver,v The desorption of gold and silver from a carbon which has been in contact with an ore pulp apparently depends upon the shifting of the equilibrium encountered between adsorption and desorption. Four methods of shifting the equilibrium succeeded in des orbing the gold and silver. The four methods comprised (1) the use of a solvent in conjunction with a large excess of precipitant in order to remove the gold from the receiving solution as the desorption progressed; (2) the use of a solvent in conjunction with electrolysis to accomplish the same purpose; (3) the use of a solvent and large volumes of solutions, added in stages, to keep the concentrations of the gold and silver in the receiving solutions sufficiently low to accomplish desorption; and (4) the use of higher temperatures by employing hot solvent solutions in a pressure chamber. The authors believe that the last method given is the most feasible, as this method is very rapid, has the lowest reagent cost, and accomplishes the desorption with the minimum volume of desorbing solution. ' 6 E. H. Crabtree, Jr., V. W. Winters, and T. G. Chapman, "Development of Carbon-Cyanidution," A.I.M.E., San Francisco meeting, 1949.

264


The treatment involved consisted of placing the charge of gold- and' silver-bearing carbon in a 10- by 12-in. horizontal batch-type pressurefilter press with filtrate valve closed, adding cyanide solution followed by hot water and subjecting the charge to full boiler steam pressure of 90 to 95 lb. per sq. in. for a soaking period of.20 min. The filtrate valve was then opened, and the pregnant solution forced through a heat exchanger and into a storage tank. Five cycles as described were used for each charge of 2.22 lb. carbon, equal to about 26 lb. carbon pel' 24 hr. (approximately the amount used to treat 7.5 to 9 tons low-grade ore). The amount of cyanide used was equivalent to 2 to 20 lb. per ton of carbon treated and the hot water to 50 tons per ton of carbon, which TABLE 37. CARBON CYANIDATION-ADSORPTION DATA (Container modification-Results of 63-hr. test at Sahuarita)

Weight, lb.

Material

Heads .................... Carbon ................... Barren solution ........... Tailing ........ , ..........

3780 8.06 6426 3780

Tons per 100 Assay, oz. tons of heads gold pel' ton

100.0 0.213 170.0 100.0

0.096 41.86* 0.0012t 0.005

Per cent

100.0 92.7 2.1 5.2

* Carbon assayed 29.7 oz. silver per ton. t Barren solution from agitator 3 averaged 0.0025 oz. gold per ton. TABLE

Material

38.

CARBON CYANIDATION-DESORPTION DATA

Carbon,

lb.

TonsperlOO tons of mill heads

Assay, oz. per ton

Gold

Silver

Per cent Gold

Silver

--- --- --Heads, carbon, screens 11 to 31 .. 8.06 Pregnant solution ...... " ........ 466 8.06 Desorbed carbon .................

0.213 12.32 0.213

41.86 0.68 0.63

29.7 0.48 Trace

100.0 98.4 1.6

100.0 100.0 0

corresponds to about 10.65 tons pregnant solution per 100 tons of ore treated. Conventional cyanide practice would require the handling and precipitating of nearly thirty times as much pregnant solution. Typical metallurgical data for the adsorbing and desorbing phases of this process are shown in Tables 37 and 38. The barren solution given in the tabulation, viz., 0.0012 oz. gold per ton, was the solution after treatment in the scavenging screen placed in the tailing launder. The barren solution from agitator 3 is given in the second footnote as assaying 0.0025 oz. gold per ton. The scavenging screen, with a very short time of contact with the tailing pulp, therefore reduced the assay of the barren solution from 0.0025 to 0.0012 oz. gold per ton.


265

Although the data show that desorbed carbon which has not been reactivated has decreased adsorptive speeds, it is entirely possible that more effective contact with such .carbon will make up to a considerable degree for the loss of adsorptive speed for such carbons. The data indicate that the reactivation of desorbed carbon restores its adsorptive speed to that of fresh carbon. The authors believe that the loss of adsorptive speed of a desorbed carbon as compared with a fresh carbon is due to the partial filling of the interstitial spaces of the carbon by slime. The reactivation of desorbed carbon, which apparently restores its adsorptive speed, is believed to increase the interstitial spaces by oxidation with steam during reactivation. Electrolytic Method. In the desorption scheme worked out at the Getchell mines the gold-bearing carbon is leached with a hot solution of sodium sulphide and caustic soda and the dissolved gold precipitated electrolytically on graphitized wood shavings using a stainless-steel screen anode. The spent electrolyte is returned continuously to the leaching step. When treating 700 tons pel' day of minus 325-mesh slimes, it is expected that the consumption of activated carbon will not exceed about 60 lb. per day. ION EXCHANGE USING SYNTHETIC RESINS

Ion Exchange. As an alternative to the use of activated carbon for the removal of gold and silver from cyanide solutions by the mechanism of adsorption, the application ofion-exchange methods using synthetic resins has been proposed. Tests along these lines carried out by the U.S. Bureau of Mines on ore from Buckhorn mining district in Nevada are described in R.I. 4374, January, 1949. The ore was chosen because its very poor settling and filtering characteristics made conventional cyanidation impractical. It was ground to minus 100 mesh for the present tests. Only the plus 35mesh fractions of the commercial resins were used, and both the anionic and cationic types were investigated. After a period of agitation with the cyanided ore pulp, the resin was separated by screening and washing on a 65-mesh sieve. The gold and silver were then removed from the resin by eluting with sodium hydroxide, after which it could be used for subsequent adsorption, though with increasing loss of capacity with each cycle. Summarizing the tests, the author states that about 78 per cent of the gold and 50 pel' cent of the silver in the ore were recovered by countercurrent adsorption and regeneration.

266


While the countercurrent method proved to be the best system, it was not practical to make use of the simultaneous dissolution and adsorption technique worked out by T. G. Chapman owing to the marked adsorption of free cyanide. by the resins. The anionic-type resins were found to be superior to the cationic type in the particular pH range and other conditions investigated, and about 25 lb. resin per ton of ore treated was required, though the actual consumption would presumably be only a small fraction of this, depending upon the undetermined amount of replacement required. A 48-hr. leach was used, followed by adsorption times as short as 15 to 30 min. for 95.4 per cent gold and 79 per cent silver removal from the pregnant solutions. Sodium hydroxide, which removed close to 99 per cent of the precious metals adsorbed, was found to be the most efficient of the eluting and regenerating agents tested. A more detailed investigation is required before the advantages and limitations of the use of ion-exchange resins, as compared with activated charcoal, can be evaluated. At the present time it would appear that charcoal is a cheaper material and one that is more resistant to mechanical abrasion and chemical deterioration," BROMOCYANIDE PROCESS

Bromo salts are a mixture of 57 per cent sodium bromide, NaBr, and 43 per cent sodium bromate, NaBr03, in the form of light-gray, lightyellow, or reddish-brown crystals or powder. The use of bromo salts for treating a telluride concentrate in the WrightHargreaves plant at Kirkland Lake is described by J. T. Willey in E. and JJ1.J., July 7, 1928. Two 12- by lO-ft. tanks were used as collectors. They had mechanical agitators and a capacity of 15 tons solids and 15 tons solution. When thoroughly mixed, samples were taken for specific gravity, moisture, alkalinity, and cyanide content. If necessary, sulphuric acid was added to reduce alkalinity to 0.1 per cent, and cyanide was added to increase the strength to 1 lb. per ton. Bromocyanide Was next made by mixing bromo salts, cyanide, and sulphuric acid. To make 1 lb. bromocyanide it requires 521 grams bromo salts in 1500 cc water, 207 grams potassium cyanide in 1500 cc water, and 486 grams 66°Be. acid in 4500 cc water. Twenty pounds bromocyanide was made for each charge of 15 tons concentrate. The procedure was as follows: 7 It is evident that with synthetic resins costing between 50 cents and several dollars per pound, the over-all consumption that could be tolerated commercially would be quite small.


267

Three 40-gal barrels were used. Into the first, which was immediately above the agitating tank, was put the requisite quantity of sulphurio acid. For 20 lb. bromocyanide this barrel held twenty times the quantity of water and acid given in the formula. The other two barrels were immediately above the acid barrel. From the bottom of each of these barrels was' a lead pipe leading to the acid barrel, which itself had a lead pipe leading to the agitator. Plugs served as valves. The acid barrel was covered to prevent escape of bromine fumes and also had a hand-worked paddle mixer. Into one of the upper barrels was placed twenty times the quantity of bromine salts and water given in the recipe, and in the other barrel an equivalent amount of cyanide and water. The'salts were then dissolved. Next, the plugs in the upper barrels were pulled out simultaneously, letting the solutions into the acid barrel, the contents of which were stirred. The 20 lb. bromocyanide was then run into the agitator. Stirring proceeded for 24 hr. when the pulp was sampled and assayed. If high in gold, more bromocyanide was run in, and the agitation repeated until extraction was complete. During July, 1927, 70 tons concentrate averaged $76 per ton before treatment and $1.50 after treatment, equivalent to 98 per cent extraction. The bromocyanide process" has been successfully applied to some mispickel-gold ores, for the reason that this process requires no oxygen for gold dissolution. Some of the reactions involved embrace the following:

+ lICNBr + lIHzO = AszOa + FeS04 + lIHBr + lIHCN + 3NaCN + CNBr = 2NaAu(CN)z + NaBr

(FeAsS)z and 2Au

The HCN is absorbed by the addition of alkali during the process. It is the opinion of Julian and Smart that the activity of the bromo salt is not due to the liberation of cyanogen, though that probably occurs, but to a liberation of oxygen according to the equation 2 BrCN

+ NaCN + 4NaOH

= 2NaBr

+ 2NaCN+ NaCNO + 2H zO + 0

Since BrCN is decomposed by alkali, it is important that free alkali should be kept very low during the treatment, and any lime necessary for settlement being added at its conclusion. BrCN

+ 2NaOH

= NaBr

+ NaCNO + H 0 2

The process was once used at Deloro mine, Ontario, Canada, for the treatment of an arsenopyrite ore and was developed in Australia by Dr. Diehl for treatment of the sulphotelluride ores of the Kalgurli district. It is practically obsolete today but should be kept in mind as a possible method for handling refractory gold-bearing materials. 8 R. J. Lemmon, "Reaction of Minerals in the Cyanidation of Gold Ores," C.E. and JJ£. Rev., March, 1940.

268

CYANIDN.rION AND CONCENTRATION OF ORES

Ammonia-cyanide Process. In C.E. and M. Rev. Dec. 11, 1939, R. J. Lemmon points out that the original work on this process was done at the University of Sidney. It was introduced into the United States in 1901 by Bertram Hunt but has been used only on one or two minor operations, since closed down, in this country. Describing this process in a later issue of the same publication (Mar. 11, 1940) Lemmon says: The employment of ammonia in excess of the amount of copper soluble in the cyanide solution, but in less amount than required to bring the total soluble copper into solution, has been found beneficial in increasing the gold solution, reducing cyanide consumption, and substantially reducing the amount of copper entering the gold precipitation unit. The mechanism of reactions includes the following: 2NaCN

+ Cu

!"

=

CU(CN)2

+ 2Na+

and 3Cu(CN)2

+ 4NI-I

10H

= 2Cu(CN)2

+ 4H 20 + [Cu(NHa),,].(CN)2

The cupriammoniurn cyanide dissociates into the complex radicle [Cu(NH a) 4] and free cyanogen [(CN)2]. A further reaction probably ensues, 4Cu(CN)2 CU(CN)2

+

(CNh,

+ Cu(NH

3 ) '1

=

4NH a·2(CuCN)2

dicuproso-cupric cyanide and cyanogen

It is believed that the (CN)2 liberated reacts with hydroxyl ions to form alkaline cyanides which then dissolve the gold. The are to be treated should have a low permanent lime alkalinity, a weak solution of cyanide then added to dissolve some of the soluble copper, followed by slight excess of ammonia, and further lime added at the end of the treatment period to precipitate the ammonia-copper solution in the are pulp. If the process has worked efficiently, the dicuproso-cuprie cyanide should also be left as an insoluble in the pulp. Should the pulp contain interferent iron salts, a small amount of a soluble lead salt may be added before the addition of the cyanide. The method is applied in a similar manner to copper "oxides" such as malachite.

Where this method was used to treat a complex gold-bearing copper silicate ore in California, the cyanide consumption was reduced from 8 lb. per ton by direct cyanidation to 1 lb. per ton. The gold was precipitated in zinc boxes by the usual method. Pretreatment to Remove Copper. Where sulphides of copper are the source of the trouble, their removal by flotation is frequently successful (the greater part of the chalcopyrite in the N oranda ore is removed by this method), but if the copper mineral is largely oxidized, chemical leaching methods may be indicated. The use of either acid leaching or ammonia leaching is suggested. The latter usually involves a higher cost, whether recovery of the copper is or


269

is not practiced, while the acid scheme suffers from the disadvantage that it is often difficult in practice to neutralize completely the residues from acid leaching and a rather carefully designed plant involving filt!-lring and agitation equipment is required to ensure successful operation (see "Calcine Treatment at Rietfontein"). . Greenawalt in Hydrometallurgy of Copper, p. 370, 1912, states that in his opinion, If the copper is extracted with an acid solution, the most logical way of recovering the gold is with chlorine, and the most logical way of extracting the silver, is with a chloride solution. The various methods of chloridizing .and of extracting silver by means of the CuCI, solution alone or by the FeCL usually present are described (see "Chlorination Processes").

Roasting Followed by Leaching. Where the above methods fail to solve the copper problem, roasting of the ore or concentrate followed by a water or weak acid leach before cyanidation is usually successful (see Chap. X, page 173). It might be mentioned in this connection that recent work at the Westport Laboratories of the Dorr Company using FluoSolids roasting technique indicates that this method can .frequently effect extremely high conversion of the copper to the water- or weak-acid-soluble form. Any such improvement in the completeness of sulphatization will be of interest to those attempting to cyanide ores carrying copper, zinc, and other basemetal values which, unless removed before cyanidation, may render the cost in cyanide prohibitive. CHLORINATION PROCESSES

In Liddell's Handbook of Nonferrous Metall1lrgy, Vol. 2, 1945, there is to be found a very complete account of the uses of chlorine as applied to the recovery of gold and silver. (pp. 336-340 and p. 524 et seq.). Introducing the chapter entitled "Chlorine Metallurgical Processes" Liddell says: Chlorine as a metallurgical agent appears to have lost ground during recent years, and much of what follows is of historical rather than operating interest. However, it is a question in the author's mind whether the great decrease in the price of chlorine and the large sources of supply that will be available after the war do not warrant reinvestigation of the applicability of the chlorine process to present-day metallurgy.

The use of chlorine can be broadly classified into (1) dry (or wet) chlorination; (2) chloridizing and amalgamation; (3) chloridizing roasting and leaching; (4) volatilization.

270

CYANIDATION AND CONCENTRATION OF ORES 1A. DRY CHLORINATION

Chlorination Process for Guld.? This process was based on the fact that chlorine, in the presence of moisture, converts gold into the trichloride Au Cl, ,.which is soluble in water and removed by washing, the gold being then precipitated by ferrous sulphate, sulphur dioxide, hydrogen sulphide, or charcoal. Coarse gold requires long contact and should be removed by amalgamation. Pyritic ore or concentrate requires a dead roast before chlorination; thoroughly oxidized ore may be treated directly. Basic ores-containing lime and especially magnesia-absorbed much chlorine and might become heated. This was checked by roasting with a high temperature at the finish to frit the magnesia with silica. Chlorination was suggested by Percy and by Plattner independently in 1848, though Plattner apparently made the first commercial application of the process to the arsenical ores of Rechenstein, Silesia. The Deetken or California process was carried out in comparatively small wooden vats with bottom filter of perforated boards, resting on slats and covered with coarse gravel and sand; a vat 8 ft. in diameter with 3-ft. staves would hold about 3 tons. The crushed and roasted ore was loosely charged by sifting in a moist condition to facilitate leaching, a cover was luted on with clay or dough, and gaseous chlorine generated in a lead vessel from manganese dioxide, salt, and sulphuric acid was admitted by a lead pipe to the bottom until it could be detected at a hole in the cover. In later practice, liquid chlorine was purchased in steel cylinders. After 12 to 36 hr. contact (adding more chlorine if necessary), water was turned in at the top, any chlorine escaping at the bottom pipe being led to another vat. The yellow solution was run to the precipitating vat, and the charge washed until the effluent was colorless. The residue was then shoveled out; if much silver was present, it was transferred to another vat and leached there with hyposulphite solution. Some ores, rich in silver, were first leached with hyposulphito and then chlorinated. The barrel process, used on a large scale in Colorado, involved the rotation of the ore in barrels of wood or heavy lead-lined steel, holding 5 to 25 tons, while chlorine was generated under pressure in the mass by means of bleaching powder and sulphuric acid. Barrels were often built with an internal filter on one side, consisting of pebbles or coarse sand confined by slotted boards and a perforated lead plate. After 3 to 6 hr. water under 20 to 40 lb. pressure was admitted by a trunnion and washed the charge in 1 to 2 hr. A ton of ore would use at least 10 lb. of bleach and 15 lb. of sulphuric acid. The ore was charged dry from a hopper and discharged by sluicing through manholes in the side. With the exception of the Malm process,"? chlorine processes for raw ores of the precious metals are necessarily confined to the treatment of surface ores or to clean gold and silver sulphide ores, in which the gold and silver minerals are not combined with the base-metal sulphide minerals.

In.

WET CHLORINATION

In the Plattner process which was in constant use from 1851 to about for the treatment of pyrite concentrates, the material after dead

1916

Liddell, Handbook of NonfB1T01ls lI1eta1l1l1"(]1I, VoL 2, pp. 336-337, 1945. The Malm process for complex ores involved dry chlorination in a three-compartment tube mill, the roasting of the tube-mill product in a multiple-hearth furnace followed by leaching in chloride solution, and the precipitation of the gold, silver, and copper with metallic lead. Zinc, lead, and other chlorides were also recovered by this process. 9

10


271

roasting was leached in water saturated with chlorine gas. In the early days the chlorine was generated by the addition of bleaching powder and sulphuric acid, but later, in the treatment of Cripple Creek ores, the chlorine was generated by electrolysis. Lead-lined chlorination barrels provided with perforated lead filters inside were used. The gold was. usually precipitated from solution by ferrous sulphate or hydrogen sulphide. 2.

CHLORIDIZING AND AMALGAMATION

The patio process for silver was invented by a miner m Pachuca, Mexico, in 1557. The dry crushed ore was moistened and mixed with copper sulphate, common salt, and mercury, the working and aging of the charge being performed in yards, or patios. The copper chlorides formed reacted with the silver to form silver chloride. The silver was soluble in the brines present during the subsequent leach, was precipitated by the mercury and immediately amalgamated. The method was not well suited to ores carrying appreciable amounts of base-metal sulphides. The pan-amalgamation process was a development of the ancient "cazo," or caldron, process. In Europe it was called "barrel amalgamation," but in the United States, where it found its greatest use, it was known as the pan-amalgamation or "Washoe process," from Washoe, Nev., where it wasB.rst operated in 1861. Liddell describes the operation as follows: The "pan" was made wholly of cast iron or with a cast-iron bottom and wooden sides. In either case .the bottom was made hollow for the introduction of steam to heat the charge. Cast-iron mulIers for grinding, stirring, and amalgamating the ore were attached to a vertical shaft in the center of each pan. The capacity of each pan ranged from 0.5 to 2 tons of ore and was usually about 1,l1 tons. The ore was first crushed by jaw crushers and then by stamps or ball mills. If crushed wet, the excess of water was removed by settling tanks. The crushed ore from the settling tanks was then shoveled into the pans. Salt and copper sulphate were added in the ratio of 5 to 10 lb. of salt and 2~~ to 5 lb. of copper sulphate per ton of ore. When the ore is free from interfering minerals, the salt has been reduced as low as 2 lb. and the copper sulphate to U
272


The chemistry of this process is the same as that of the patio process, except that the iron of the pan and mullers also acts as a reducing agent, not only for precipitating silver in metallic state from its chlorides but also for preventing the formation of any chlorides or sulphides of mercury, and in this manner avoids the chemical loss of mercury mentioned under the patio process. The ore must be siliceous or neutral in character to avoid precipitation and the loss of effective copper salts by the carbonates of lime and magnesia, although the iron of the pan has a tendency to reduce these salts to copper and hence militate against their effectiveness.

3.

CHLORIDIZING ROASTING AND LEACHING

11

The gradual exhaustion of oxidized ore and the increase of base-metal sulphide minerals with the silver sulphide minerals, together with increased facilities for transporting fuel and supplies, led to the introduction of chloridizing roasting and the attendant leaching processes. In chloridizing roasting, salt in amount equal to 5 to 15 per cent of the weight of the ore is added for a sulphur content of 2.5 to 3.0 per cent, which, if deficient, is made up by the addition of pyrites. The Wedge-type multiple-hearth furnace was most frequently used, and the temperature was not allowed to exceed about GOODC. The Patern process makes use of the solubility of silver chloride in sodium'thiosulphate solution. Following chloridizing roasting and a warm water wash, the ore is leached with Na 2S20a·5H20 and the silver precipitated from the purified effluent with sodium sulphide. The Augustine process consisted in giving a chloridizing roast to silver ores or matte and leaching with a strong solution of common salt, which dissolves silver chloride to a limited extent. The silver was then precipitated by metallic copper, and the copper by scrap iron. The cement copper was then melted and granulated for reuse. The Tainton process included the scheme of chloridizing roasting of oxidized silver lead ores and leaching with spent chloride electrolyte containing free chlorine and ferric chloride, to which bleaching powder had been added. The solution was then electrolyzed in specially designed closed cells using rotary cathodes, and the silver and lead deposit collected and melted to bullion. An alternative scheme was to deposit the silver and gold preferntially by using a lower current density. The chlorine was used to generate the hyprochlorite above mentioned as well as the calcium chloride needed to control the S04 concentration of the solutions. The Holt-Dern process was developed at Park City, Utah, and reached oornmercial operation in 1941 when a plant was constructed at Silver City by the Tintic Milling Co. The process consisted of a chloridizing roast in a Hclt-Dern furnace, followed by a percolation leach with a nearly saturated solution of common salt acidified with sulphuric acid. The silver was recovered by precipitation on de tinned scrap iron.

A modification of the Holt-Darn process was used at the Achotla chloridizing mill of the Cia Minera de Penoles in Guerrero, Mexico." Here 11 12

Ibid. H. P. Allen, "The Aehotla Chloridizing," T.P. 773, A.I.M.E., February, 1937.


273

the mixture of oxide ore, sulphide are, and salt (3.3 per cent of the charge) was given a chloridizing roast in 16 Holt-Dern blast roasters.. The hot calcine was transferred to leaching vats and first given a water wash, which was drained to storage for subsequent precipitation of gold and silver values on scrap iron. . A weak acid wash, derived from spraying the flue gases, was next applied, and the effluent passed over scrap-iron trays. Following prolonged washing to neutral reaction, the are was cyanided, using solutions carrying 1 kilogram per metric ton KCN and 0.3 to 0.7 kilogram CaO. The mill had a capacity of about 6500 tons per month, and about 90 pel' cent of the gold and 84 per cent of the silver in the o~'e were recovered by the process described.

4. VOLATILIZATION 13 Loss of gold by volatilization was recognized from time to time and was the cause of serious monetary loss in treating gold-silver ores by chloridizing roasting, but the reason for this loss seemed to be little understood. In 1891-1893 Croasdale discovered that a commercially complete (above 90 per cent) volatilization of gold could be obtained from Cripple Creek ores by roasting with salt. About the same time, Pohle independently obtained similar results with silver ores from Aspen, Colo. Systematic investigation of the volatilization of metals as chlorides was begun by these men in 1898 and was carried on with a large-scale experimental plant until 1903. Numerous.investigators have worked on this process since that time. By raising the temperature of a chloridizing roast to 1050°0. it has been found that gold, silver, copper, and lead can be commercially volatilized as chlorides from their ores and the metals recovered from the fumes. By charging the 01'6, salt, and sulphur mixture as quickly as possible into the hot zone of the furnace, volatilization begins at about 750°0. and is completed within 30 to 60 min. The roasting atmosphere must be kept highly oxidizing. The process is continuous, and the ore is commercially devoid of value as it discharges from the furnace. The furnace used for this process was a regular cement kiln, 100 to 125 ft. in length, fired in the usual manner. Gold is easily volatilized, but in what form was never definitely determined. It was generally supposed that gold trichloride was formed at low temperatures, and this was decomposed into metallic gold and chlorine at temperatures below 300°0. If this is true, the metallic gold formed from the vapors is probably colloidal and is carried out of the furnace in that form with the gases. Rose states that, when gold is heated in chlorine at atmospheric pressure, trichloride of gold is formed, which volatilizes at all temperatures above 180°0. up to and beyond 1100°0. Other metallurgists thought that gold forms a double chloride with salt or other metallic chlorides and volatilizes in that form. The theory of a double chloride seems more probable. Silver is less easily chloridized and volatilized than any of the common metals. It seems to be extremely sensitive to atmospheric conditions in the furnace and may be affected by the gangue constituents in the ore. Silver chloride melts to a thin 13

Liddell, op. cit.

274


liquid at about 451°C. before it volatilizes, which probably accounts for its sluggish volatilization. It is much more easily volatilized in the presence of other metallic chlorides, which would indicate that it volatilizes as a double chloride. The metallic chlorides are driven from the ore in the form of vapor, and they condense as cclloidal particles of fume. Cottrell precipitation is required for complete recovery. The metals are recovered from the collected fume by substitution of one metal for another in aqueous solution or by electrolysis of the fused chlorides.

Articles in C.E. and M. Rev. for Feb. 10 and Mar. 10, 1943, describe laboratory volatilization tests on two different products: The first was a hematitic calcine residue carrying 7.2 dwt. Au per ton. Salt (5}!i" per cent by weight) was added, and the change nodulized and thoroughly dried. Using a rotary kiln externally heated, with the outlet gases passing over hot charcoal, a 25-hr. run at 900 to 1000°C. resulted in 86.8 per cent of the gold being volatilized and 95.9 per cent of the volatilized gold recovered on the charcoal. It was stated that sufficient gas flow must be used to remove products of combustion or any reducing gas that would tend to break up the auric chloride. Lake View and Star flotation concentrates were used in the second case. B. H. Moore found that, by using 10 per cent salt and heating in a closed muffle at 800°C. for 40 min., 92.9 per cent of the total gold was volatilized. The author believes that a closed furnace is essential for best results because of the rapid decomposition of AuCIa in air. Furthermore, there is a much smaller volume of gas to be handled than in an air-swept furnace and a better chance to collect the fine gold particles, which in the laboratory was done in a water column using glass beads.

Recent work at Columbia University'< on the rate of solution of gold leaf showed that the addition of sodium chloride to chlorine or bromine solutions accelerates the dissolution of pure gold and of (95.8 per cent) gold leaf immersed in these solutions while sulphates and fluorides have practically no effect. The conclusions reached were that the accelerating effect of chloride ion has not been recognized heretofore but that, if it is assumed that the first reaction between chlorine and gold isI 5 2Au

+

Cb - 2AuCl

and that the rate of solution of the gold is controlled by the rate at which the insoluble aurous chloride film is removed from the surface by the reactions AuCl

+ Ch + Cl-

and AuCl

+ or =

AuCh

it is believed that the results are satisfactorily explained. 14 G. L. Putnam, "Chlorine as a Solvent in Gold Hydrometallurgy," E. andM.J'J Vol. 145, No.3, p. 70. 15 Compare this reaction with the Elsner equation for solution of gold in cyanide,

p.211.


275

It was shown that aurous chloride is insoluble in water but is soluble in sodium chloride. The aurous chloride is later oxidized to the auric state, but in dilute chlorine solutions this is not the rate-controlling reaction. THE HANDLING OF WASTE CYANIDE

SOLUTIO~S

Tailings pulp carrying traces of cyanide and the discard of barren solutions in some cases constitutes a hazard to both humans and animals," and methods have been devised for destroying the contained cyanide. A recent paper "The Treatment of Cyanide Wastes by Chlorination" by J. G. Dobson" published in the Sewage Works J., November, 1947, discusses the subject as follows: oJ Free cyanide is one of the most toxic components of industrial wastes that are often discharged into sewers and streams. Because its toxicity to fish varies with pH, concentrations of other ions present, temperatures, and oxygen content of the receiving water-as well as activity of development and species 'of fauna in the steam-it is difficult to set up exact limits of concentration which can be safely discharged. Under some conditions, as little as 0.1 p.p.m. has proved fatal to fish. From 0.1 to 0.3 gram of CN is fatal to humans. For 35 years or more various investigators have sought a satisfactory method of treating cyanide wastes. Among the methods that have been used are acidification and removal of the resulting HCN gas by air blowing, reaction with "lime sulphur," aeration, treatment with ferrous sulfate, and oxidation with potassium permanganate.

The first of these methods is, of course, the standard scheme used for cyanide regeneration already described; the second involves the addition of commercial lime-sulphur solution, whereby the CN is converted to the cyanate. A definite yellow precipitate indicates complete reaction. The third (aeration) has been discussed by C. H. Clevenger and H. Morgan in "Atmospheric Decomposition of Cyanide Solutions," p. 413, Mining and Scientific Press, 1916; the fourth, or ferrous sulphate method, is described by J. Moil' and J. Gray in "The Destruction of Cyanide" J. C..M.S. S. A., Vol. 10, p. 433, 1909; Vol. 11, p. 152, 1910; and the fifth, oxidation with permanganate, is discussed by E. F. Eldridge in "Reducing 16 An amusing sidelight on this subject and one that is illustrative of some of the social problems confronting mill operators in the early days of the West was the socalled profession of "metallurgical cattle raising." This practice, like that of "smoke farming" with which the smelter people had to contend, netted certain irresponsible characters a considerable profit at the expense of the mining companies. Just as crop failures from a number of causes were conveniently blamed on the presence of fume from the smelter stacks, so the loss of a cow that died from drinking water draining from the tailings ponds was blamed squarely on the men who operated the cyanide plant, and the courts usually awarded the damages claimed. 17 Presented at the 22d Annual Conference of the Michigan Sewage Works Association, Jackson, Mich., JVlay 16, 1947.

276


the Toxicity of Cyanide 'Waste" Eng. News-Record, Vol. 111, No. 23, p. 677, Dec. 7, 1933. All the above methods have been used commercially for the reduction of the cyanide content of concentrated solutions, but they all leave a substantial cyanide residual. This residuai may be serious unless extremely large dilutions are available. Another objection to all these methods is that the pollution load is increased by the addition of objectionable chemicals. None of these treatment methods are applicable to the treatment of dilute wash solutions such as are obtained from plating operations. Chlorination of solutions at pH above 8.5 has now proved to be an economical and satisfactory method for treatment.

Referring to the chemical reactions involved, Dobson says: When chlorine is added to a free cyanide solution with sufficient free alkali present to maintain a pH above 8.5, the cyanide is oxidized to cyanate: NaCN

+

2NaOH

+

Cl,

->

NaCNO

+

2NaCl

+

H 20

This reaction is practically instantaneous. Under all conditions which we have tested, it is complete in less than 1 min. Theoretically, it requires 2.73 parts of chlorine per part of CN and 3.08 parts of caustic per part of CN. However, most trade wastes contain a substantial portion of the required free alkali. If an excess of chlorine is added, free chlorine residual, as measured by the O-T-A test, will be found in the solution after 1 or 2 min. In all trade wastes and laboratory samples that we have tested, no cyanide within the limits of titration accuracy has been found in the presence of free available chlorine at pH 8.5.

Part II Descriptive

1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 1 . 1 1 1 1 1 1 1 1 1 1

I

CHAPTER XV

Treatment of Gold Ores This chapter is devoted to brief descriptions of the current methods of treating simple and complex gold ores containing relatively small amounts of silver. Examples are taken from the principal gold-prod1lcing centers of the world, and the subject matter is claseified. under specific types of treatment schemes. These vary from simple amalgamation "10 the use of more elaborate methods involving gravity concentration, flotation, roasting, and cyanidation. CLASSIFICATION OF MILL FLOW SHEETS

The various treatment schemes we find in use today for the extraction of gold and silver from their ores have been evolved as the result of a gradual improvement in metallurgical techniques over many years. The degree of complexity of the flow sheet used in any specific case will usually depend upon the refractoriness of the ore itself, but it must also be kept in mind that in many instances the kind of equipment and type of treatment used is not necessarily the most modern practice for the type of ore handled. Older methods are likely to be retained unless the life of the mine, cost of operation, and percentage recovery justify replacing the older equipment. Since the economics and techniques of ore treatment are in a continuous state of flux, mill flow sheets are continually undergoing change, and unless the mill has been built in the last few years, it is usual to find some parts of the operation more up to date than other parts and much old equipment still in use simply because it does not pay to replace it. An attempt is made here to classify mill flow sheets into five main types, with subdivisions depending upon whether or not roasting is used. These types are arranged in order from the simplest recovery scheme for the handling of free-milling gold ores to the kind of complex flow sheet such as might be required where the gold is intimately associated with base metal and other sulphides. Making due allowance for minor variations in practice, it is believed that this general classification scheme will apply to milling operations throughout the world. SYSTEM OF CLASSIFYING FLOW SHEETS

Type I. Amalgamation (with or without corduroys, jigs, or traps). Type II. Straight cyanidation (with or without Type I steps). Type III. Cyanidation followed by concentration (with smelting or cyanidation of concentrates). 279

280


Type IV. Concentration with treatment of concentrates. Type V. Concentration with smelting or cyanidation of concentrates and cyanidation of tailings. Type II is subdivided into (a) all-sliming plants, (ss) separate sand-slime treatment, (1') roasting before cyanidation. Types III, IV, and V may also utilize corduroys, jigs, or traps and are subdivided into (e) oyunidation of raw concentrates, (1') roasting before cyanidation, (S) smelting of concentrates.

Type I flow sheet is not found on a large scale today. Its use is limited to small-scale prospecting and placer operations, where it appeals to the operator of small mining claims because of its extreme simplicity. By itself, however, this treatment scheme is seldom capable of extracting a high percentage of the values from the average gold ore.. Type II is the most common flow sheet in use today for the extraction of gold and silver from comparatively simple ores where special treatment of some refractory portion of the ore, e.g., finely disseminated gold in sulphides, is not necessary. The greater number of the mills in Canada and South Africa employ this flow sheet. The all-sliming plant with grinding in cyanide solution tends to replace the older method of grinding in water and separate treatment of sands and slime. It is rare nowadays to find plants where the whole ore is roasted before cyanidation, although this used to be common in Australia and is practiced in one plant in the United States today. Type III is a type of flow sheet that has evolved from the search for methods to reduce cyanide-plant residues. Frequently the gold that fails to dissolve is closely associated with sulphides that may be concentrated by flotation, and this small amount of refractory material reground and given special cyanide treatment, with or without roasting. Alternatively, the recovery of base metals in this step may justify shipping the whole product to a smelter where both the gold and other metals are recovered. Further discussion of the advantages and disadvantages of this scheme as compared with Type IV flow sheet will be found on page 129. Type IV. Where technically feasible, this is the most economical type of flow sheet, in that the cyanidation capacity required is only a fraction of the tonnage milled and intensive treatment, in regard to either fineness of grind or time of contact, is usually possible at a relatively low treatment cost. As pointed out in Chap. IX, the whole success of this scheme depends upon how low grade a tailing can be made in the concentration step (usually flotation). Type V. This treatment scheme is used where a certain gold-bearing fraction of the ore, which can be separated by gravity concentration or other

TREATMENT OF GOLD ORES

281

means, requires fine grinding and intensive cyanide treatment or even roasting followed by cyanidation. In this case, however, in co~trast to Type IV above, it is not feasible to make a final tailing in the concentration step, and cyanidation of the remainder of the ore is also necessary. This is also the flow sheet commonly used for the treatment of base-metal ores carrying gold, where the concentrates are smelted and cyanidation is really a scavenger operation to extract what gold remains in the flotation tailings.

SECTION I. NORTH AMERICA CANADIAN PRACTICE !

Gold is mined and milled in five principal districts in northern Ontario, in four districts in northeastern Quebec, in Manitoba, and in British Columbia. Table 39 lists the larger operating mines in these districts together with type of flow sheet and 1946 gold recovery and costs in dollars per ton, the latter figures being taken from the Financial Post of Jan. 29, 1948. There are also a number of base-metal producers in various parts of Canada where the recovery of gold and silver is an important factor in the total output, but these are not covered in the present list. ORE TREATMENT IN NORTHERN ONTARIO

Types of Ore Deposits. P. E. Hopkins of the Provincial Department of Mines in Ontario Gold Deposits, Their Character, Distribution, and Productiveness, 2d ed., 1924, shows 24 lode gold areas, all in Pre-Cambrian rocks. To the end of 1934 these had produced 25,240,000 oz. gold with a peak production rate in 1939 of over 3,000,000 oz. per year. Ontario contributes nearly 60 per cent of the total Canadian output. The pyritegold-quartz type of deposit, in which the pyrite is predominant, is the most cornmon and most productive. It is represented by the Porcupine deposits. The gold-telluride veins, characterized by the presence of tellurides, as at Kirkland Lake, are another prominent type of deposit. Kirkland Lake Ore and Its Treatment. Ore from the mines of Kirkland Lake may be generalized as a silicified altered porphyry containing gold, free and in pyrite; tellurides (principally altaite, a nonauriferous lead mineral); chalcopyrite; and other important minerals. Some of the gold exists as a telluride, but Kirkland Lake ore as a whole is not a telluride ore, although some from the Lake Shore mine is considered as such. The tellurides are found throughout the ore zone and average less than 0.1 per cent. Pyrite amounts to 2 per cent or less. The ore is hard and requires fine grinding to liberate the gold. Gold extraction will average over 95 per cent. The Kirkland Lake Gold Area, Vol. 36, Part II, 1928,

282


TABLE

39.

PRINCIPAL CANADIAN OPERATIONS EMPLOYING CYANIDATION,

Mining area

Mine

Porcupine, Ontario

Hollinger Me Intyre-Porcupine Dome Mines Buffalo Ankerite Coniaurum Paymaster Consolidated Hallnor Pamour Preston East Dome

Kirkland-Larder Lake, Ontario

Lake Shore Wright-Hargreaves Kerr Addison Chesterville Teck-Hughes Sylvanite Upper Canada Kirkland Lake Macassa Mines

Mutachawan, Ontario

Matachuwan Consol. Young Davidson

Little Long Lac, Ontario

Little Long Lac MacLeod-Cockshutt Hard Rock

Red Lake, Onta-. rio

Cochenour-Willans MeKenzie Red Lake Madsen Red Lake Pickle Crow

N oranda-Rouyn, Quebec

Noranda Stadacona

Cadillac-Malartic, Quebec

Consol. Central Cadillac Canadian' Malartic Sladen Malartic East Malartic

Lamaque, Quebec

Golden Manitou Lamaque Sigma Siscoe East Sullivan (building)

Rated capacity, tons per day

Flow sheet type

Recovery, dollars per ton

1946 Cast, dollars per ton

--- --- --5300 Vc 8.71 7.63 IV c 10.79 2500 8.18 1700 IIa 9.50 7.02 1300 Vc 5.93 6.47 600 IIa 9.09 7.89 600 IIIc 8.54 7.70 Vc 17.23 10.86 400 IV c 3.41 1600 3.10 IIa 7.95 1000 7.35 --------2700 III r 15.05 11.40 1200 III c 18.73 11.94 2250 IIa 7.25 4.88 IIa 900 4.57 0.96 600 IIa 11.97 11.24 600 IIa 10.96 8.54 IIa 12.19 350 9.49 400 IIa 11.97 11.10 IIa 15.19 11.74 400 -----4.01 1000 IV c 3.29 1050 Vc 3.21 2.67 --------325 III r 10.74 10.19 700 IV r 7.86 10.69 450 IVr 6.98 --------170 IV r 13.81 12.09 8.60 8.61 250 IIa 400 IIa 9.97 7.62 480 IIa 18.99 14.93 --------3600 IV sc IIa 6.47 5.95 475 -----400 IIa 1000 IIa 4.09 3.60 700 IIa 3.36 3.82 1600 IIa 5.25 4.96 --------7.19 1000 IV c 1200 IIa 9.34 7.62 4.82 1200 IIa 6.15 1000 IIa 5.46 5.66 2000

283

TREATMENT OF GOLD ORES TABLE

39.

1946

PRINCIPAL CANADIAN OPERATIONS EMPLOYING CYANIDATION

(Conlimted)

Mining urea

, Rated capacity, tons per day

Mine

Flow sheet type ---

Duparquet, Quebec

Beattie Consolidated

1200

Manitoba

Hudson Bay San Antonio Gold Mines

6000 400

British Columbia

Yellowknifc, Northwest Territory

.,

Cost, dollars per-ton

--- --2.92 8.31

--- --Vs 15.63

9.83

I1a

------

Pioneer Cariboo Gold Quartz Island Mountain Hedley Mascot Kclowna Exploration Silbak Premier

350 350 150 225 275 550

Giant Yellowknife Negus Mines Con-Rycon

500 190 300

I

IV r

Recovery, dollars per ton

I1a I1a I1a Vs III s IV s

14.22 16.98 10.11

15.37 14.49 8.89

9.77 12.70 -----IV r

III r III r

by E. W. Todd, of the Ontario Department of Mines, gives details of the ore deposits. Porcupine Ore and Its Treatment. Ore from the mines of the Porcupine district may be generalized as quartz and mineralized schist, carrying from very little to a considerable amount of free gold, and gold in pyrite, which ranges from 2 to 8 per cent. There are other minerals, but they are of little significance, except pyrrhotite in the Dome. Treatment of these ores varies, as will be noted in the descriptions of representative plants. Other Producing Areas. In addition to the considerable gold-mining industry in British Columbia, which ranks third in production in Canada, mention should also be made of the important developments in the Yellowknife district of the Northwest Territories, where rich gold strikes were made some years ago. A brief description of one of the mills operating in this area is given on page 307. Canadian Production." After an uninterrupted decline since 1941, Canadian gold production increased slightly in 1946 compared with 1945. Silver output, however, which had decreased annually since 1940, eontinued its downward trend, reaching 12,676,928 fine oz. in 1946, the lowest since 1906. Output of gold in 1945 and 1946 distributed by provinces is shown in Table 40. 1

Minerals Year-Book, p. 115, 1945.

284

CYANID.ATION AND. CONCENTRATION OF ORES

Porcupine United (Type I). As far as is known, there are no straight amalgamation plants of any size or importance in operation today. As a matter of historical and technical interest, however, a brief description of the old Porcupine United mill at Timmons, Ontario, follows. The ore consists of quartz with fine stringers of schist, of banded quartz schist, or of stringers of quartz in schist. About 75 pel' cent of the gold is free and fairly coarse, according to R. A. Vary in I.e. 6433, U.S.B. of M., 1931. Half-inch ore is ground to 65 pel' ceut through 200 mesh in a 4~2-ft. by 16-in. Hardinge mill. Amalgamation, as indicated in Fig. 61, is done as follows: The ball-mill discharge falls into a distributing box where water is added, and the flow distributed to two 4- by 8-ft. copper plates ~& in. thick and with a slope of 1% in. per ft. The box traps considerable coarse gold. It is cleaned out once a week, and its contents run through an amalgamation barrel. The table frames supTABLE

40.

CANADIAN GOLD OUTPUT BY PROVINCES 1945

1940

Fine oz.

Fine oz.

Alberta ......................... British Columbia ............... Manitoba ....................... Northwest Territories ........... N ova Scotia .................... Ontario ......................... Quebec ......................... Saskatchewan. ................. yukon ......................... .

7 186,854 70,655 8,655 3,291 1,625,368 661,608 108,568 31,721

105 123,348 78,732 19,738 4,579 1,835,887 586,231 112,000 47,023

TotaL ............ '" .........

2,696,727

2,807,643

Province

porting the amalgamation plates are constructed of 2- by 4-in. lumber, with crosspieces placed on edge and spaced 6 in. apart. The table decks are made ,of I-in. shiplap on which copper plates are screwed. The slope of the deck can be changed to suit conditions by placing wedges between the deck and the table frame. The plates are not silvered and have to be treated with a weak cyanide solution before the mercury will amalgamate with the copper. Little trouble has been experienced' by copper showing on the plates. Mercury is shaken on the top half of the plates, and none is added elsewhere. Ordinarily the plates are dressed every 3 hr., but oftener if the ore is rich. The method employed in cleaning a plate is to by-pass all the feed to the other plate, clean off all ore particles, then brush the top of the plate. The loosened amalgam is removed, and if this leaves the plate too dry, mercury is shaken on and rubbed in well. The plate is then brushed horizontally, working from the center to the sides and starting at the bottom and working to the top of the plate. Any amalgam or Idose mercury adhering to the sides of the plate is then brushed to the top or removed if the amount is appreciable. On the morning shift the plates are given an extra brushing, and mercury is added to loosen the amalgam. Then the amalgam is stripped off with a

285


piece of rubber conveyor beIting, stripping being done at right angles to the slope of the plate; the amalgam is lifted; and the plates redresse I in the usual manner. Care is taken that the plates are not stripped too clean.

:6" t

'MIi'1 Ore

.4 ~"x

Hrdinge Mill

,

2-4 xB Amalfmatlon plates ~------ 2-42"x4B" Blanket slril2.s

t

-

~-------- Amalrram tral2.s

-"-r-

J

~-------- Bucket elevator

t

Do" classifier

-,

f

(Overflow) ~-----

(Sands)

l

Gibson amalf].amator

t -

James Table (Cone)

7

"""(TailS)

L

L T o waste,'

Shippedto cyanide plant

Clean-up----------:l

t

16 " diam x 36" amalgam barrel

t

Collectinf]. box

t-

r - Amalrram I2.late

,-"'-,

'Amalgam press

Ret~ted ..

(Tails)

l__-i"~mill Return to circuit

Melting furnace

t

Gold bullion FIG. 61. Flow sheet of the Porcupine United mill, Ontario, Canada. Little crystallization of the copper tak~s place, and the plates are rubbed occasionally with a weak cyanide solution, which removes any tarnish or stains. Enough water is used to maintain an even flow of pulp over the plates, and when the ball-mill discharges too much coarse material, the feed is cut off for a short period. Forty per cent of the total gold recovery is made on the plates.

286


On the lower end of each table below the plate is fastened a sheet of }&-in. iron plate, 42 by 48 in. A blunket of No.6 silence cloth is bid on this plate and secured by a flat iron bar }& in. thick and 2 in. wide bid on top of the blanket. The bar is held by notches cut in the table frame. The pulp from the amalgamation plates passes over the blankets; pyrite, fine gold, and mercury from the plates are caught on the blankets. The blankets are changed and washed in a tub after each dressing of the plates, the blanket concentrates being sent to an amalgamation barrel for treatment. The plate and blanket tailings drop to amalgam traps at the end of each table and are elevated therefrom by a bucket elevator which returns them to the Dorr classifier. This elevator has a deep sump which is a good trap. The sump is cleaned out at regular intervals. Amalgamation takes place in the whole circuit. Elevator discharge launders, classifier and other launders all collect rich sand and amalgam; these are cleaned out periodically and treated in the amalgamation barrel. Thirty-five per cent of the gold is recovered by barrel treatment. The rake product from the Dorr classifier is returned to the ball mill, and the overflow runs by gravity to a Gibson impact amalgamator, attached to a James sandconcentrating table. The amalgamator catches float mercury and fine gold which has escaped the amalgamation plates, blankets, and traps. The amalgamator is opened and washed, and its plates are scraped once a week. The James table is operated at 250 r.p.m. with a %-in. stroke. The table concentrates, averaging $40 per ton, are dewatcred and sent for cyanide treatment to an affiliated company. The table tailings are elevated by a bucket elevator to the tailings dump. The table concontratos and table tailings are both sampled hourly by hand by taking dip samples from their launders. The concentrates or sands from the blankets, trups, and launders are ground for 10 hr. in a cast-iron amalgamation barrel, 16 in. in diameter and 36 in. long, revolving at 22 r.p.m., using worn balls from the ball mill as grinding media. Then about 250 oz. mercury and 3 lb. slaked lime are added to the charge, and it is again ground for 5 to 8 hr. The barrel is washed out into a box, the iron balls are carefully cleaned by hand, and the residue is run over a small amalgamation plate to the mill circuit. The mercury and amalgam are collected, washed, and cleaned with hot water and then squeezed by hand through fine sheeting to eliminate excess mercury, retaining the amalgam in the form of a ball. The amalgam is retorted outside the mill over a wood fire at regular intervals, using a cast-iron retort which has a capacity of 1000 oz. The sponge-gold recovery is 35 to 40 per cent the weight of the amalgam retorted, and the mercury loss is small. The sponge gold is melted in an oil-burning furnace at the affiliated company's refinery. Soda, borax, and manganese dioxide are used for flux, and the molds are coated with lampblack. The bullion is sampled by drilling small holes in opposite ends of the bar at top and bottom. The average grade of bullion is 770 fine in gold. and 120 to 140 fine in silver. . The crew for the 25-ton mill consists of two amalgamators working 8 hr. each. On the day shift the mill is operated by a mill foreman who takes care of cleaning the amalgam, retorting, and melting. ·The ore averaged $11, the concentrator heads or copper plate tailings $2.80, the concentrator tailings $1.80, and the concentrates $40 per ton.

Preston East Dome Mines, Ltd. (Type IIa). This is one of the more recent all-cyanidation operations in the Timmons district, the original mill of 500 tons' capacity having been started up in 1939 and subsequently

287


increased to 1000 tons' daily capacity. The following account is taken from "Milling at Preston East Dome" by R. D. Lord, C.M.J., August,

1941. 3-1,000-ton steel mill bins 6'x 14' Allis chalmers btl mIll

Barren solution

+---1

2- 8'x 60" Harding,e boll mIlls

..

3-16"X24"Duiex Denverjig,s

--

3 -60"high-weir Akins classifiers (Satds)

0"rJ.. canrntrale

(Ove'ftlow)

-r

2-6"x 30"x 12' Dorr bowl classifiers

r-

't'

(Sands)

4'x6' boll mIll

f.

(Overflow)

4-30'Hardin{J..e trax thickeners

t

(Underflow)

-

t, .

HY..draulic cone

t

-,

(Overtlow)

4 -20 'x 2 4 Agjlators

-r--

3'x4'AmO'lJ.a'tn barrel

-.-,,-r

f

(Overflow) (Underflow)

~

2-12 x 14 Drum filters - f - - - - - - , .

'f Rel2ul{J.ers

--r-

Amolgpm 5el2.0ra;dtor

Filtrate

to.soutia:

~/gam I2.lale

storage

Retort

2-20'x24' Agitators

tGold

,

2-12 x 14 Drumfilters-----..I

'f

9'x 10' Rectangular clarifier

Re'P..ul.'P..ers

-l-

t

4 x 10 Crowe vacllllm tonk

Tailing,s to waste

" t

2-36 IT-frame I2.resses Filtrate Barren solution storag!!..

precf'late

-,

Meltin{J.. furnace Gold

FIG. 62. Flow sheet of the Preston East Dome mill, Ontario, Canada.

The ore is a porphyry with scattered mineralization consisting mainly of 1.8 to 2.0 per cent pyrite, though some pyrrhotite is present. The average gold value is about 0.22 oz. per ton with about 10: 1 gold-silver ratio.

288


All the ore, with the exception of a small proportion hoisted from brow bins, is passed through 12-in. grid grizzlies underground. The ore carries 3.5 to 6 per cent water, mostly absorbed in handling. Crushisu; equipment consists of an 18- by 36-in. Traylor jaw crusher, followed by double-deck Niagara screens, the oversize from the top deck passing to a standard Symons cone crusher and the oversize from the bottom deck, which is fitted with a ~~- by l-in. ton cap screen, passing to a short-head Symons. The product of both cone crushers falls onto a common cross belt and is recirculated through the screens. The screen undersize constitutes the mill feed and passes by belt conveyor to the 3000-ton mill bins. Grinding is carried out in a circuit comprising Allis-Chalmers and Hardinge ball mills closed-circuited with Akins classifiers. The overflow from the latter pass to Dorr bowl classifiers which overflow a final product at about 70 per cent minus 200 mesh. The bowl sands are returned to the mills. An interesting feature of this mill is the installation of Denver mineral jigs to catch coarse gold as it is released by grinding. A 16- by 24-in. duplex unit is placed in the grinding circuit between each ball mill and primary classifier. Nearly 60 per cent of the total gold recovery in this mill is made in these jigs. The hutch product is treated after regrinding by amalgamation. For details see Chap. IX. Thickening before agitation is carried out in Hardinge tray thickeners with the trays connected in parallel. Maximum settling rate is obtained at a lime concentration of 0.60 lb. per ton CaO, but variations in rate occur due to varying amounts of sericitic material in the ore and to seasonal temperature changes. Pulp is drawn from the thickeners at 47 per cent moisture, and agitation is carried out at this dilution, since filter capacity hardly permits a lighter feed. Agitation. A contact time of 35 hr. is provided in the four primary and two secondary Hardinge agitators. There is a filtration step between the two stages of agitation, using barren-solution washes and repulping. Following final filtration on two more of the 12- by 14-ft. low-submergence drum filters, the cake is repulped in water and pumped to the tailings pond. Sodium cyanide is fed into the system at the primary-classifier overflow launders, and strength in the agitators maintained at 0.4 lb. per ton of solution. Lime strength averages 0.3 lb per ton CaO. Further operating data are shown in Table 41. An over-all gold recovery of 97 .4 per cent is made in this mill. Dome Mines (Type lIa). We are indebted to the management for the following details of operations at the Dome mill. Since its discovery in 1909, the property has passed through a period when it was considered


289

that the life of the mine was over, has twice survived- the calamity of having its milling plant completely destroyed by fire, and is now one of the most important gold producers in Canada. Metallurgical problems have been very serious at times, but these have all been overcome, with the result that the mill is now making exceptionally fine metallurgical extractions. Blanket practice has been developed to' a high state of efficiency in the present mill. The description of this part of the flow sheet is given in Chap. IX. Figure 63 shows the flow sheet of the crushing plant and mill. The run-of-mine ore is crushed in a Farrell jaw crusher to 4-in. size and further reduced in a 5~-ft. standard Symons cone crusfier which is set for a %-in. product. It may be worthy of note to mention at this point an innovation introduced into the crushing plant a few years ago to avoid trouble with the Symons crusher, the vibrating screens, and the rolls due to wet fines contained in the ore as it comes from the mine. The feed to the vibrating TABLE

41.

STEEL AND REAGENT CONSUMPTION

Grinding balls (3 in.) Lime , , Cyanide , Mercury Lead acetate Zinc dust

1.3 lb. per ton milled 1.2 lb. per ton milled 0.4 lb. per ton milled 0.0036 lb. per ton milled 0.006 lb. per ton solution 0.33 lb. per ton solution

screen following the primary jaw crusher and ahead of the Symons crusher is washed copiously with water, and the minus ~:4:-in. material along with the wash water from this screen is dropped into an Akins classifier. The overflow from the classifier is pumped direct to the mill surge tank, while the sands from the classifier join the dry-crushed ore for delivery to the mill bin by conveyor. The Symons product is passed over Hummer vibrating screens with ;!-s- by %-in. openings. The oversize goes to 18- by 42-in. rolls set at 7:4: in. and this product returns to the Hummer screens. The undersize from these screens is delivered to the mill bin by a belt conveyor. Ore from the mill bin is fed by conveyors to three 8-ft. by 30-in. Hardinge ball mills, using 4-in. balls. The discharge from these goes to five duplex D 6- by 18-ft. 4-in. DOlT classifiers, which in turn feed five 5- by 22-ft. tube mills using No.4 Danish pebbles. The product from the tube mills and the overflow from the classifiers are pumped to the blanket plant. The tailings from the blanket plant are pumped to two large cones, which distribute the flow to 16 smaller cone classifiers. The spigot discharge from these cones is pumped to four duplex C Dorr classifiers. The rake

290


Mine are

,~

Farrell jaw crusher

~

-t

---,--20:"--rVibrating screen

x v4"openif/[l

(Undersize onawater)

f

t

Akins classifier

(Overflow)

l

(Oversize)

.

.

t (Rake product)

Pump

l

Surge lank (Oversize)

t

I "/'1-_"1 5Y.fI1ons cone crusher 5 !2.....L!
Hummer screens(3/8" X

%"op.enings)

+ - ,- '-t=:':'::-:::'--o-"t , versue -,--

1,800-lon sloragebm

..

3-Hardinge ball mills 8 'x30 "

~

18x42 " rolls

l

5-00rr dUp'lex classifiers 6'x18~4"

E:.~'E....!..!!!j!~;==-__2Q--r--=~=erflow 4-00rr Ihicken),s 40 'x 14 '-4

tt

~waler

10 mill

/d!!!§ -----~. 4-Pachuca aflalors 42 x 14' {;ypnide - - - - l..~ 4-Pachuca alYalOrS 42 xl4'

,--r, 2-Surge lanks

t Fillrale 10Wecip.ilalion

ttl

5-Memll p'resses90 -4 x6-6 frames

t '[gjj§.,

FIG. 63. Flow sheet of the Dome Mines mill, Ontario, Canada.


291

product is fed to two tube mills, 60 in. by 22 ft. inside dimensions, which use 174:-in. steel balls. The overflow from the model C Dorr Classifiers is returned to the primary, or model D, classifiers, and the product from the regrinding tube mills joins the flow from the primary grirrding units and is pumped to the blanket plant. The overflow from the 16 cone classifiers flows to four 40-ft.-diameter by 14-ft. 4-in. Dorr tray thickeners. The clear-water overflow returns to the general mill circuit, while the underflow is pumped to four 42-ft.diameter by 14-ft. Pachuca agitators where lime is added, and the pulp is agitated for about 77~ hr. This preliminary agitation increases cyanide extraction and reduces cyanide consumption. The aerated pulp then has cyanide solution added to it and is pumped to another series of Pachuca tanks, where it receives about 117~ hr. agitation in cyanide solution. The pulp then discharges into two storage or surge tanks, which in turn feed by gravity to the five Merrill filters with ninety 4-in. frames each. The unclarified solution from these presses is then pumped to a Merrill simultaneous leaf clarification and Crowe precipitation unit. The pulp, discharged from the Merrill slime press, is thickened in Dorr thickeners and pumped into the tailings-storage dam, the thickener overflow being used for sluicing out the Merrill presses. In the report to shareholders for the year ended Dec. 31, 1947, the following results of the mill operations for the year are given: heads, 5.4873 dwt, per ton; recovery, 96.14 per cent. This is an increase from 95.08 per cent at the beginning of the period. The consumption figures for the year ending Dec. 31, 1947, were (in pounds per ton milled) cyanide, N aCN, 0.36; lime, 2.31; zinc dust, 0.067;' and lead acetate, 0.013. Kerr-Addison Gold Mines, Ltd. (Type IIa). The mill at this property, which is situated in McGarry Township, Ontario, about 26 miles east of Kirkland Lake, is an excellent example of the more modern, allcyanidation flow sheet. The complete operation from mine ore to gold bullion is shown in Fig. 64, which represents one of the three treatment units and one of the two precipitation units. Two kinds of ore are being mined, one a green carbonate carrying only a trace of pyrite, the other a silicified flow carrying up to 5 per cent pyrite. Average content of mill feed is about 1 per cent pyrite. The ore carries about 0.2 oz. silver. per ton with a 20: 1 gold-silver ratio. The mine rock is reduced in jaw crushers underground to 4 to 6 in. before hoisting to the surface, then fed to a 57~-ft. standard Symons crusher. The discharge goes to vibrating screens fitted with J-in. square mesh screens. The undersize is conveyed to the mill bins, while the oversize is returned

292


Minerocf<

r-

2-36" x 48" underground jaw crusher

t t

6-lon mineSkil2s

-

600-lon are bin

t

-.-

Jeffrey.feeder

»2

:==J

ff. slondord Symonscone crushers

+

rty. ! t/

3-4'xI0'Sim

5ueer

.

(Undersize) (O/erSiz/f)

3-1,o00-ton fine are bins

Grind solulion lank

-

t Barren solulion 10

f

8'xl2'Rodmill

I

mills andclassifiers

I

l>

7-D.5.F.X. Dorr classifiers

l (Overflow)

B'x f5' Ball mill

l~ (Overflow)

[ 8':15' Ball mill

t

! j

Overflow '1',. tank Pregnant SOlution

3-36' 3-Comp't Iray thlckeners--l Lime and Cyamde

L

:J

t (Rake)

24 'X12'Dorr bowlclassier

l (Rake)

shari-head Sy'mons

-

t

100-Leaf clarifier

9- Darr agiTators in senes

,

t

NaCN~

znfeeder-j

2-C.C.P. washing, stageswith 3 - 36 3- campt. Dorr tray thickeners in eachstage

4-Precil2itatian I2resses

4-11~' xl6'Oliver filtersl

Gold bunoo

--. + t Rel2ull2ers

Tailmgs dam

Filtrate to Grind solution storage

T

3-Wobi furnaces

t

Barren solution tank

Overflow pari to pregnant tank, mast to Grind tank

FIG. 64. Flow sheet of the Kerr-Addison mill, Ontario, Canada.


293

to a 5~~-ft. shorthead Symons and then goes back to the screens. This is typical of the modern crushing circuits used in Canada. There are three treatment units and two precipitation units, but since they are almost identical, only one of each will be described. Each treatment unit handles 1400 tons of ore per day, and each precipitation unit, 5500 tons of solution per day. . The I-in. ore is fed directly to an 8- by 12-ft. rod mill, using 37~-in. rods. The product discharges to a 7-ft. Don classifier in closed circuit with an 8- by 15,-ft. ball mill supplied with 28 per cent 3-in. and 72 per cent 2~~-in. balls. The overflow from this classifier is pumped to a 24-ft. Don bowl classifier in closed circuit with an 8- by 15-ft. ball mill with 1~-in. balls. The discharge of this mill and the overflow of the 7-ft. classifier are pumped together to the bowl by one 6-in. sand pump (and aspare), There are thus three stages of grinding. Grinding is done to 93 per cent minus 200 mesh in cyanide solution. The bowl overflow at 3.5 to 1 dilution is pumped with a 6-in. sand pump (and a spare) and split three ways to three 36-ft., three-compartment Don tray thickeners, which make an overflow that passes to precipitation and an underflow at 50 per cent solids that is pumped by quadruplex Don diaphragm pumps to a series of nine DOlT agitators, which give a 40-h1'. period of contact. Lime is added to the bowl classifier and the last agitator discharge, and cyanide is added to the first and fifth agitators. Washing is accomplished by two stages of C.C.D., with three 36-ft., three-compartment tray thickeners in parallel in each stage, and by one stage of filtration in four 117~- by 16-ft. Oliver filters, with flood washing. The repulped cake is pumped to a tailings-disposal area. Pregnant solution from the primary thickeners is clarified in a 100-leaf clarifier and passed through Crowe vacuum towers. Zinc dust is then added along with a drip of strong cyanide solution ahead of the precipitation presses. It has been found necessary to treat this cyanide with lead acetate in order to remove soluble sulphides, because they cause trouble in precipitation. About 0.5 ton of barren solution per ton of ore milled is discarded to control the concentration of fouling agents, principally KCNS and nickel, which give trouble in precipitation. A cleanup is made once every 2 weeks, and the precipitate melted in three Wabifurnaces. Sillimanite linings are used, and the method adopted is to produce a clean, high-grade bullion along with a high-grade slag. The slag is remelted in a Carbofrax-lined furnace to produce matte which is shipped and a low-grade slag which is added to one of the rod mills. Further data on this mill are shown in Table 42. The over-all gold extraction is 97 to 98 per cent, and the cost about 72

294


cents per ton milled, of which crushing averages about 12 cents per ton and milling 50 cents per ton. Cariboo Gold Quartz Mining Co., Ltd. (Type IIa). This 375-ton mill located at Wells, British Columbia, Canada is of special interest because of the parallel C.C.D. and stage filtration circuits employed. The C.C.D. section was installed in the original mill, but when increased capacity was required, the filtration system rather than a second C.C.D. unit was installed because of the saving in floor space and heating requirements in winter. The gold occurs in the ore both as free gold in quartz and as free gold on pyrite surfaces, though some is intimately associated with the pyrite. Galena and sphalerite occur in small amounts, and cosalite, a sulphobismutite, is frequently encountered. The sulphides comprise 15 to 20 per cent of the ore, which is soft and grinds easily. The ore is fed through a grizzly to a coarse ore bin, thence to a 3- by 5-ft. Ty-rock screen with H-in. openings. The undersize passes directly to the TABLE

42.


Steel consumption...... .. ... . . . . . ... . .. 2.70 Cyanide consumption. . . . . . . . . . . . . . . . .. 0.80 Lime (burned) " .. . . . . . . .. 2.10 Lead salts '" 50 Cyanide strength.... . . . . . . . . . . . . . . . . . . . 1.00 Cyanide strength.... . . . . . . . . . . . . . . . . . .. 0.40 Protective alkalinity. . . . . . . . . . . . . . . . . .. 0.40

lb. lb. lb. lb. lb. lb. lb.

per per per per per per per

ton ton ton ton ton ton ton

milled milled milled milled solution in agitators solution in precipitation solution

fine ore bin, while the oversize is crushed in a lO- by 20-in. Traylor primary crusher and a 28-in. Traylor gyratory secondary crusher. Both machines are protected by a 20-in.-diameter Dings magnet and a Dings magnetic pulley. The primary grinding circuit comprises a 54H-in. Marcy mill in closed circuit with a 4- by 20-ft. Dorr classifier with restricted-type overflow. The secondary circuit makes use of a 5- by 21-ft. 8-in. Dorr classifier. The secondary classifier overflows at 15.8 per cent solids a pulp ground to 54 per cent minus 200 mesh. This pulp is split in the proportions shown on the flow sheet and fed to the two cyanide circuits, which consist of Circuit 1: one primary 20- by lO-ft. Dorr tray thickener, followed by four Dorr agitators giving 33 hr. retention at 50 per cent solids, countercurrent washing in three more thickeners of the same size, and a final washing of the last underflow on a lO- by 5-ft. Dorrco filter. Circuit 2: one primary 24- by IS-ft. Dorr tray thickener, followed by four agitators giving 28 hr. retention at 50 per cent solids, washing in two stages on ll-ft. 5-in. by 14-ft. Oliver filters with a tray thickener between the washing steps.

295


I

Grizzly

2

Coarse ore bin

3

Vibrating screen

4 5 6

Jawcrusher

Minerunore

Magnels

Gyratory crusher

7

Ane ore bin

8 9 10 II

Primary bol/mill Primary classifier Secondary classifier Se<;ondary bolfml1f

/2

Troy thickener

13 /4

Clorifier Vacuum tank

/5

Precipitation press

/6 Furnace '17' Agll%r 18 Filler 19 Repulper 20 Barren solution storage

21

GrInding solu/ion s/oroge

Q402 OZ Au,

+---------

_=~~d'-8

f) 21

'01

#/' ,

~'

I I I

0015solulion

I

. Q042 ~Qlufion.-.J

----

50%soliciSl ---~~Wo/er ____-..J QOO6 solution

19 Unciisa 0194 Diss. :0007 Tolol ,0201

TAILING TOWAS7E Uncilss, 0,0198oz.Au/lon

~7gi g.8f8g~~~:

FIG. 65. Flow sheet of the Cariboo Gold Quartz Mining Co. mill, British Columbia; Canada.

296


Combined pregnant solution from both circuits is clarified on a leaf clarifier and precipitated in a Merrill-Crewe system. The Shrivel' presses used are protected by an electric eye. The solution and pulp values are shown at each stage in Fig. 65. The milling cost for 1941 was $1.12 pel' ton, a figure that includes $0.31 for power which is generated by Diesel engines on the property. Ball and reagent consumption for the same year are shown in Table 43. The over-all gold recovery for the year 1941 was 94.87 per cent. The solution strength in the grinding circuit was 0.96 lb. NaCN, 0.55 lb. CaO pel' ton of solution, and 0.95 lb. NaCN, 0.91 lb. CaO pel' ton of solution in the final agitator discharge. TABLE

43.


Ball consumption Cyanide (NaCN) Lime Zinc dust Lead salts TABLE

44.

2.85 lb. per ton milled 0.625 1.513 0.127 0.085 GOLD DISTRIBUTION IN TAILINGS

Per cent distribution

Mesh

Per cent

Gold assay

Mesh value

+65 +100 +150 +200 -200 Sands -200 Slimes

1.65 8.57 13.45 15.50 6.98

0.010 0.005 0.012 0.030 0.029

0.0002 0.0004 0.0016 0.0046 0.0020

1.0 2.0 8.2 23.6 10.3

53.85

0.020

0.0107

54.9

0.0195

100.0

100.00

Wright-Hargreaves (Type IIIe). A very complete account of milling operations at this property is to be found in a series of articles entitled "Wright-Hargreaves Improves Its Milling Practice" by Malcolm Black, mill superintendent, published in E. and lYI.J. for March, April, and May, 1939. The flow sheet today is substantially as described with the exception of certain changes made necessary partly by reduced tonnage and partly by improvements in milling practice during the past 10 years. The management has kindly supplied the necessary information to bring this account up to date. Operations were begun in 1921 with a l50-ton mill which was enlarged to an 800-ton unit during the period 1921-1929. A straight cyanide circuit was employed, but a froth rich in gold tellurides was removed in


297

2-IJOOO-ton are bins

-~

...---.. Allis Chalmers Boll mill 9'x 7'

·l Dorr F. t:s~x 30' is:4-Bowl clas-;;;J;;;8' ix: Diam. ~ Double-scoop- feeder X.

L-

...

.........

(Ttfh

w)

Lime

zt:

. 4,.A.C. ;;;;r;;,ilIS 5'x 16'

re;------~~;2?:J.-D2!o'2!.rrr..!LthljJ,ic:.&keners 50'diam.

'"'-------l~ -~t 2-Dorr anitJorS' 24' x 26'

NaCN

I

.

J

------J.~

~~w~

t

Dorr ag/fator 22'x20' -- ~ .

Melting furnace

2- Oliver filters 14'x16'

. 70 solution storage

t

--rRep-ulp-ers

.

/

40' Southwestern air-flotation cells

. t

. ,

u

. 2-Banks of 7-56 Fagerg.ren flat. cleanercells ~

(Conc)

Do~-" thicken:r 35' x 9 '-4"

~

Ballmill}X22' .. -.--. 3.-Denver-Wallace agitators

.

- - 70 Lake Shore mill

t

Oliver filter //'..6"x 14'

t.

(Tails)-----,t

Lime'

retreatmentplant See fig. 65

I) .NaCN

~.

FIG. 66. Flow sheet of the Wright-Hargreaves mill, Ontario, Canada.

298


K &; K flotation machines and treated by the bromocyanide process (see Chap. XIV). The present mill with a daily capacity of 1200 tons was completed in 1933. With the finer grind now used, the nominal rate is not greatly in excess of 800 tons per day which, owing to the present labor shortage, has been temporarily reduced to 500 tons per day. The present flow sheet comprises two-stage grinding in an Allis-Chalmers ball mill closed-circuited with a quadruplex DOlT classifier (only one-half of which is currently used) and bowl classifiers closed-circuited with AllisChalmers tube mills. The grind averages about 85 per cent minus 325

FIG. 67. Twelve-foot-wide DOlT FX classifier in closed circuit with 9- by 7-ft. AllisChalmers ball mill at Wright-Hargreaves, illustrating raising of ball-mill discharge to classifier by means of spiral-scoop elevator and gravity return of classifier sands.

mesh, and the overflow from the bowl classifiers at 10 per cent solids flows to two DOlT thickeners. The underflow, at about 55 per cent solids, is pumped from each machine by a Denver Quad diaphragm pump, discharging into the primary agitators (two DOlT machines 24 by 26 ft. and in series). The portion of the effluent from the Dorr thickeners required for precipitation flows by gravity to clarifying leaf filters, and the remainder is sent to mill storage. The overflow from the primary agitator 2 flows by gravity to two 14- by . 16-ft. Oliver filters. A barren wash is used. The cake drops directly into a repulper of the Dye-Davis type, repulping in barren solution. The pulp is next pumped at 50 per cent solids by means of a 5-in. Allis-Chalmers

299


S.R.L. pump to a 22- by 20-ft. Dorr agitator which has been converted to act as an agitator, and no further dilution is made. The pulp next passes to the second stage of filtration, which, like the first, consists of two 14- by 16-ft. Oliver filters followed by repulpers. 'Water is used on thet.wo front sprays, and barren solution on the others. The sprays on this bank are fitted so that all or any number may be used for either water or solution. The repulped filter cake then passes to a flotation circuit consisting of Southwestern roughers and Fagergren cleaners. A small amount of pine oil only is now used, the object being merely to float a rich froth which is given a prolonged cyanide treatment after thickening and filtration in a series of three Denver-Wallace agitators. Lime at the rate of 25 lb. per ton of concentrate is added to the regrind, or mixing, mill just before the pulp passes to the agitators, where strong N aC~ solution is added to bring the pulp dilution to about 2 to 1. Of the gold present 85 to 90 per cent is extracted. An interesting point is that, after this agitation in the retreatment circuit, the residues are returned to the primary agitators and no building up of values takes place provided that not all the sulphides are floated. It is noteworthy that in the flotation circuit not more than 0.5 per cent of the total sulphides contained in the flotation heads is floated, and the average pyrite content of concentrates is about 7percent FeS2. The average value of the ore milled for the first 5 months of 1947 was close to $10 (at $20.67 per ounce), and the over-all gold recovery 97.1 per cent. Primary agitator feed carried about $1.50, the Southwestern cell feed $0.40, and flotation tails $0.321. Other data are as follows: TABLE

45.

GENERAL MILL DATA AND COSTS

Item

Mill data

Precipitation ratio ........... Pregnant solution ............ Reagents: Cyanide (KCN) ............ Lime .......................

3.S2:1 $4.33

Crushing Ball milling

0.69 lb. per ton ore 0.50S lb. per ton are

Tube milling All other

Zinc ....................... Steel (ball mills) ............. Steel (tube mills) ............ Hp. load per ton milled per day ........................ Power cost ...................

0.69S oz. per ton sol. 3.109 lb. per ton 1.633 lb. per ton 3.0S $2S.53 per hp. per year

* Flotation reagents $O.OOS per ton milled.

Process

Total

Mill costs per ton milled

$0.109 0.307 0.216 0.S30* $1.462

300


This mill is of special interest because of a number of improved methods of control, which will be found discussed in the chapter "Plant Control." The mill tailings are pumped to a steel standpipe 30-in. in diameter by 89 ft. high above ground level. A 12-in. wood-stave pipe connected near the bottom conveys the tailings to disposal. At the time of writing these tailings were being custom treated at an adjoining property for recovery of the small amount of gold locked in the sulphides. Kelowna Exploration Co. Mill at Hedley, British Columbia (Type Ills). This 350-ton mill is of particular interest because a flow sheet

FIG. 68. Large DOlT bowl classifier at Wright-Hurgrcaves in closed circuit with secondary mill, overflowing at 98 to 99 per cent minus 200 mesh.

has been devised which involves the separate reduction and treatment of the hard and soft constituents of the ore, thereby overcoming a serious primary slime problem which at one time threatened to close the mill down. The principal ore mineral is arsenopyrite 'with lesser amounts of pyrrhotite and chalcopyrite. The gold occurs as extremely fine particles locked in the arsenopyrite. The ore comes from two sources: That from one ore body is very hard, provides grinding media for the pebble mills, and is ideal stamp-mill feed; that from the other ore body is very soft, and its chloropal content neces-


301

sitates the separation of primary slime from sands for successful cyanide treatment. The dry crushing, as shown in Fig. 69, is carried out in such a manner" that 4- to 5-in. pebbles can be passed to a storage bin for the pebble mills after primary breaking in a Traylor jaw crusher. The minus 7~-in. ma:-. terial carrying a large part of the primary fines is separated from the minus j:~-in. stamp-mill feed and is ground in a separate ball-mill classifier circuit. The product is treated with 0.15 lb. pel' ton lime, passed to two 30- by 12·:.ft. thickeners, and the clear overflow discarded. The stamp-mill product discharging through ~16-in. battery screens at 10 per cent solids passes first to three DOlT and one Akins classifier where the ore is deslimed and the slime passed at 2 to 4 pel' cent solids to the primary slime thickeners above mentioned. The sands pass to a tube mill in open circuit with a Dorr bowl classifier that is, in turn, closed-circuited with a second pebble mill. The bowl overflow then passes at 13 per cent solids to a separate Dorr Torq dewatering thickener, and the overflow either discarded or reused as required. The ore has now been ground in a water circuit to 84 to 85 per cent minus 325 mesh and is ready for cyanide treatment, which is carried out in two separate circuits as shown. Following agitation at 43 per cent solids, the primary slime is thickened and washed on an ll-ft. 6-in. by 12-in.F.E., Inc., string filter using flood washing. This type of filter was found to be superior to the conventional drum type on this particular problem. The primary slime-free pulp is thickened after agitation and filtered in two stages on 14-"by 14-ft. Oliver drum filters. The usual barren wash is used on the first stage, and water washing on the second stage. Pregnant solutions from both sections pass to a common tank and are clarified and precipitated in the conventional manner. The filter cakes from both sections are repulped, conditioned in two stages, and floated to recover the arsenopyrite (see Chap. VI for details of this flow sheet). The concentrate, which represents a ratio of concentration of about 14: 1, carries 38 per cent arsenic and 1 oz. gold per ton. The over-all gold recovery in this plant is 93.5 per cent, of which 71 per cent is by cyanidation and the remainder in the flotation concentrate. Solution strength in both circuits is maintained at 1.0 lb. KCN and 0.10 lb. lime per ton of solution going to the agitators. Copper sulphate, 0.40 lb. per ton; reagent 301, 0.15 lb. per ton; and pine oil, 0.10 lb. per ton, are used in the flotation, which involves a 2-hr. conditioning and a 20-min. flotation period. Both the gold precipitate and the arsenopyrite concentrate are shipped to a smelter for recovery of gold bullion.

302

CYANIDATION AND CONCEN'l'RATION OF ORES

300-lon Mineore bins

t. T '/4-.!!p..enll7Q.. 4" x 8 Ty.'ock screen , t 4 sid SY..mons cone crusher 24 " x 36 tr Troyjor JOW crusher 3 "

3 " -l4

' - -

"~

f6'X9' DenverLJil/on screen...,

Boll ml'll bin

g,lOO-lon botlery" feed bin

I

-~

.

t

-

" l

4 -5 pebbles 10 storaqe bin for pebble mt'lls I"

n'4

t

f

6'x6'A.C. boll ml'l/] a-Botteries of1,050 Ib slomp..s

4 ~

-21 DOff clossifier

1

rLtine

, , tf } 30 x 12 DOffthickener 35 'x 12'Doff thickener

. }

g~~ide

t

45"~kIl7S

I

(Sohas)

t

I

4-5'x22'p..ebblemills

t

Overflow to woste

--Dorrogilolor

Clossfier

(Overflow)

-

~

~,.

,,,.t

3 - 4-6 x 22 DOff clOSSI,Ier

I

8'X30'XI0'DOffbOWIClosslfie~

zx:

L(over.~/OW)

.

~

5'x22'f2ebble mt'll

48'xI2' DOff lorq. Ihickener Lime. } cy ol7lde r3-30' x16' Devereux ogjtolors

.

BOffe.n sotation}

2-30 'x 12' DOff thickener ~

t

11~6"xI2'F.E Inc.Str. fI'llerl

t t

G=t

3 -16 'x 14' DOff ogjlolors

t -

48'x12' DOff thickener

t

14'x 14'otver filler

Pregnont solution Clarifier lank.

t

Mem'l/-Crowe J2recig,i lion

t

Gold PPTT to smeller

Regplg,er -114'x 14' NorthfOundrY.. f,'ller

[ 1 8 'x 12' condilianer_ Soda ash ~ Capper sulphate 8 xanthate 6'x6' condifioner.K.~--"'--::r~--------' vonc.. . ~ff!!l(en flat. cells~PlI7e 011

t

t

t

6-No. 15Denvercleanercells

2 raws of 6-36 Fag§!Jj(en cells

f

28 'x 10' DOff thickener-.4'x6' Oliver f!'ller

f

t

lilt'lings to wasle

Concentrate to smelter

FIG. 69. Flow sheet of the Kelowna Exploration Co. mill at Hedley, British Columbia.


303

Lake Shore Mines, Ltd. (Type IIIr). The Lake Shore mill, the largest in the Kirkland Lake district, was originally an all-sliming straight-cyanidation plant, but for a period running through 1934 flotation of the cyanidation tailing was adopted with separate additional cyanide treatment of the concentrate. As a result of experimental and research work beginning in 1933 and extending through 1935, flotation and separate' treatment of the sulphides were discontinued. At the beginning of 1936 straight cyanidation with improved chemical treatment was established and rendered combined cyanidation and flotation unnecessary.. but finer grinding and increased time of agitation proved to be desirable, and flotation tests were conducted with the sole object of floating sulphides enclosing gold which could be treated only by roasting and reeyanidation. Shortly before the war a sulphide-treatment plant was put into operation. The results of the experimental and research work carried out at Lake Shore, including detailed descriptions of methods used, test data, and the 1936 flow sheet, are fully described in a paper "Milling Investigations into the Ore as Occurring at the Lake Shore Mine" by the staff and published by the Canadian Institute of Mining and Metallurgy, 1936. A second paper, "Fine Grinding Investigations at Lake Shore Mines," Trans. 43, C.I.1I1. and M., 299-434, was published in 1940, covering 7 years' experimental work on the grinding of Lake Shore ores. This paper includes a detailed discussion of the special methods devised for subsieve sizing control using the Haultain infrasizer, A third paper, "Roasting and Flotation Practice in the Lake Shore Mines Sulphide Treatment Plant" by J. E. Williamson in collaboration with A. L. Bloomfield (managing director) and B. S. Crocker (mill superintendent), covers the latest metallurgical development that has resulted in a substantial increase in the over-all gold recovery from the refractory portion of the Lake Shore are (see Chap. X). The tonnage rate has been considerably reduced in recent years owing both to shortage of underground labor and to conditions peculiar to the mine itself: Originally designed to handle 2600 tons per day, it was operating at about half capacity in 1948. Partly on this account and partly owing to improvements in grinding technique, the crushing and grinding flow sheets have been undergoing maj or changes. At present five stages of series grinding are used in reducing the ore from % in. to 90 per cent minus 325 mesh, though the best arrangement for mills and classifiers is still being studied. First-stage grinding is done in rod mills to about 6 mesh; with the mill closed-circuited with Simplex classifiers, following which there are four stages of tube-mill grinding using ball

CJ.:)

o

H'-

a

~..... ti

~o

Z

"..

Z

ti

a

o

Z

a

t'I

~ ~

l:3.....

o

Z

o

"'J

o

~

t'I

U1

FIG. 70. Lake Shore Mines, Ltd. This famous property is one of the larger gold producers in the western hemisphere, with profits at one time attaining 8 million dollars per year.


305

charges down to 72: in. in the last stages. Bowl classifiers receive the discharge at each stage and overflow a final product, while the sands pass to the succeeding grinding stage." The mill pulp is first treated inpreliming agitators at 10 percent solids for 47~ hr. contact time; here the protective alkalinity is maintained "a.t 1.0 lb. CaO per ton of solution. This step facilitates the dissolution of the tellurides and improves subsequent settling in the thickeners, from which barren solution is taken off for clarification and precipitation. Approximately 3.5 tons of solution is precipitated per ton of ore treated. Secondary agitation is carried outin seven Dorr agitators at 45 pel' cent solids for a period of about 48 111'., after which th~ pulp is filtered on drum filters using three sprays of barren solution and two water sprays for flood washing. The cake is about % in. thick, and a very high displacement of gold solution is obtained. The filtrate returns to the grinding circuit. Protective alkalinity in the agitators is maintained at 1.8 lb. CaO per ton of solution. TABLE 46. FLOTATION REAGENTS USED AT LAKE SHORE Reagent

Lb. per ton

H,SO" (as SO,) ......... , ...

11 to 13 0.10 0.025 0.10

oeso,......................

Pine oil ................ , " . Reagent 301 (xanthate) .....

Where added

Spray tower First conditioner Part to conditioner, part to cells To surge tanks ahead of cells

The sulphide retreatment starts at this point, the first step being to repulp the filter cake and pass it into a series of three surge tanks, the first one of which is used for mixing in customs tailings from two other mills in the district. The next step is to readjust the pulp to a pH of approximately 7.0 by contact with S02 gas from the roasters in a specially designed spray chamber (see J. E. Williamson report referred to on page 159), condition with suitable reagents, and float the sulphide content of the ore. The reagents used are given in Table 46. The difficult nature of the flotation problem is evident from the fact that there is only about 17~ per cent sulphides present in the ore and the feed to flotation is virtually all minus 325 in size (40 per cent minus 10 microns). Yet in spite of this and also of the presence of slimed gangue material, the sulphur content of the flotation tailings or final mill tailing is reduced to about 0.095 per cent. It was found necessary to thicken the flotation concentrates to 63 per cent solids in two 30-ft. superthickeners in order to get efficient filtering of the 2 See Addendum Report to "Fine Grinding Investigations at Lake Shore Mines," Trans. 44, C./.M. and M., 379-395, 1941.

306


~

&J2. from

Lime

5-8%+325mesh:' QrindinQ mJ7ls 23·5% +28 Microns

t 30'x24' I---~--i~~ 3-Primorx...J!gJlators

t

"

3 - Dorr traxthickeners 50 x12 - ,

-t t

t

NaCN

4-Dorr agitators 30'x 24'

I I--~==:::::

3-AQitators 30'x24'

Lime

•

Solution to clarification and precipitation

t

~

,t:teltinQ furnace

r

Filler SurQe tank40 x 13-4

f

5-0liver ftYters 14'x 14' Re.et.eers

[

-1-, ,

--.--

Filtrate to grinding

ci~cui;

2-SurQe tank 30 x 24

-.

Constant head tank

---r-

Stack ~",o... s_--i"'. S02 absor.etlOn tower- Excess nos to stack ~gJlator

"

and conditioninQ tanks

f..

3- Banks of B-56 Fagpg!en flotation cells

t

-

(cooc.)

f

(70115)

L....7o waste

Bank 8-56"Fal1ergten flotation cells

t-

f

(T0I7s)

(conc.)

2-su.eer-thicteners 28'x 12' -~ 2-0liver tittere 8'xIO'

. t

10-Sl2tndle Edwards roasters

-

t Rel2u!t2.ers ...

-t-

Ball m/'l/3'x6 '

~Ta waste Lime

I

(f------

NaCN

4-~ 12'X8'...-J Oliver filler !'X6'

~ Filtrate

1 t

Cake Pulpers

1'----i.. .

7O waste

FIG. 71. Flow sheet of the Lake Shore mill, Ontario, Canada.


307

extremely fine sulphides. A Sullivan slapper is used on the filter in oi'der to reduce the moisture content to a point where the cake can be handled on a belt conveyor and delivered to the Edwards roasting furnaces (see Chap. X, "Roasting and Calcine Treatment," for details of this' operation). The discharged calcine is broken up in a 3- by 5-ft. ball mill-and cyanided at 41 per cent solids in four batch agitators, followed by filtration on an Oliver filter. The filtrate is returned to the secondary agitators, and the cake repulped and sent to waste. The over-all recovery of gold in this mill is approximately 97.5 per cent, of which 96 per cent is obtained in the main cyanide circuit. It will be seen from this that the retreatment plant must, of necessity, operate on a narrow economic margin. The high efficiency obtained and low cost of operating this section of the mill are of special interest. Negus Mines, Ltd. (Yellowknife) (Type IIIr). The mill is now treating about 185 to 190 tons per day of 0.5-oz. gold ore. The ore is equally divided between two primary grinding closed circuits, one using a 5:!/z by 6 Allis-Chalmers mill in closed circuit with a 30-in. H W Akins classifier and the other a Marcy 6- by 4~~-ft. mill in closed circuit with a 4-ft. SFH Don' classifier. Grinding is to 80 pel' cent minus 200 mesh at 20 per cent solids. Classifier overflows go to three 9 by 5 tables, then to a 30- by 14-ft. ATB thickener. Flow is split to two agitation circuits: one with three 16 by 18-ft. Don' type A's and the other three 12- by 12-ft. Don' type A's. The overflow from each agitator circuit is thickened in a 22- by lO-ft. ATB thickener, the underflow from each being separately filtered on 8- by 8-ft. Olivers, Thickener overflow goes to precipitation, and cyanide tails after filtration are floated in a bank of four No. 15 and six No. 12 Denver cells. Float. tails are filtered and then go to tailings. Float concentrate is thickened in an 8~ by lO-ft. Denver thickener, then filtered on a 4- by 4-ft. Oliver. Filter cake is stockpiled for possible future roasting. The cyanide circuit is presently extracting about 75 per cent of the gold. Cyanide tailings, which are floated, average about 0.12 oz. per ton. Flotation picks up about 65 per cent of the remaining gold, the float concentrate averaging 1.2 oz. per ton and the float tails about 0.04 oz. per ton. Ratio of concentration is about 20: 1. McIntyre Porcupine Mill (Type IVc). The McIntyre mill in the Porcupine district presents a somewhat different type of flow sheet from that of Hollinger, in that a finished tailing is made by flotation concentration. This operation has been very completely described in a paper by J. J. Denny in the November, 1933, issue of E. and. MJ. and in a paper by

308


P. D. P. Hamilton, Trans. 112, A.I.M.E., 630. The following is taken in part from these publications and in part from information supplied through the courtesy of the management to bring the description up to date as of July, 1948. McIntyre ore consists of quartz, porphyry, schistose basalt, and dacite with a pyrite content of 3 to 15 per cent. The gold is associated with both the quartz and the sulphides, predominating in the latter. A typical analysis of the ore fed to the mill over a period of 1 month is given in Table 5 in Chap. III. The mill at present has a capacity of 2400 tons. The general scheme of treatment consists of crushing the run-of-mine ore in jaw and cone crushers to ~ in. and reducing the crusher product to ~{6 in. by rolls in closed circuit with vibrating screens. The screen product is ground to 8 per cent plus 65 mesh in tube mills operating in closed circuit with unit flotation cells and classifiers. Classifier overflow is floated, with elimination of a finished tailing. Concentrate from the flotation cells is reground in cyanide solution to minus 325 mesh in tube mills, operating in closed circuit with classifiers. Next comes agitation of the pulp and thickening, followed by three stages of filtering and washing, with agitation between stages. Final residue goes to waste. Precipitation of the pregnant solution is done by the Merrill:..Crowe system. Grinding and Concentration. Figure 72 gives the flow sheet of the grinding and flotation plant. The fine-crusher product is ground in single stage in five Allis-Chalmers 5- by 16-ft. tube mills. Each mill is operated in closed circuit with a single No. 500 Denver Sub-A flotation cell and a 6- by 30-ft. Dorr duplex classifier. The mill discharges are fitted with 3mesh screens; the oversize goes directly to the classifier, and the undersize to the cells and then to the classifiers. About 75 pel' cent of the gold is recovered in the unit cells, of which 60 per cent is floated and 15 per cent removed every 24 hr. from a cone on the bottom of the cell. These unit cells are built with a small hydraulic cone in the bottom to trap gold that is too coarse to float. This prevents a building up of values in the tube-mill circuit. Hydraulic water added at the bottom dilutes the pulp in the cell and cleans the concentrate therein. Concentrates removed from the cone are added to a concentrate-regrinding tube mill without any detrimental effect. The "primary" flotation section consists of 48 No. 24 Denver Sub-A primary cells arranged in eight units of six cells each. The primary flotation tailings go to eight similar banks of "secondary" cells. About 98.7 per cent of the pyrite is recovered in the concentrate. The concentration ratio is about 8.5: 1. In the flotation circuit, the pH varies from 8.9 in the tube-mill discharge to 8.4 in the secondary-cell discharge. About 50 tons tailings pel' 24 In.. is pumped to a 15-ft. Wilfley table and

309


-4,000-ton fmill are bm S _Allls _Cholmr s tubemills SxI6' S - Oenver unit cells No.SOO -------i~

.

t

s-oon dug/ex classifiers 30'x6'

[

.x:

(son~

I

8- units of 6 Denver No. 24 sub-A flotation cells inl2arallel

-L

7iJlls

•

.Distributor

C:nc.)

----i~

8- Units of 6 Denver No. 24 sub-A flotation cells in parallel

L (Conc.)-----l~

(7iJ)ings)

7iJ waste

..,J

oorrclassifier 4'-6" x 14'

l

(Sands)

Tube mill closed- circuited with oorr classifier

l

80wl classifier *+-----....

t

t

(Sands) (Ove~flow)

I

Americanfilter 8'-6 "

I

Tube millclosed circuited with classifier .

l t , Tray- thickener SO dlam. ~ t (Underflow) (Overflow) (Overflow)

4 -Agitators

t

Filtrate -+- Water to mill head tank

4-Agitators 80wl clastfler 20' diam. Sands

~

----.---t, ----.--

Oorr trax thickener SO'x 14' - - - - - . . . , 2-American filters 8~6" 8 diSC.

Oorr agitator 24 x 20

~

Overflowto clarification andprecipitation

t

r

Meltingfurnace

2-American filters 8'- 6" 8 disc.

l

Dorr agitator 24' x 20'

~

2-Amerlcan filters 8'-6" 8 disc.

t

Filtrate to lsi bowl classifier dilution

Tails to waste

.

*Raking mechanism removed

FIG. 72. Flow sheet of the McIntyre Porcupine mill, Ontario, Canada.

310


serves as a visual guide to the flotation operators. Flotation promoters are added to the tube-mill circuit and frother at both the tube-mill circuit and primary and secondary circuits. Reagent consumption, in pounds per ton of ore, is reagent 301, 0.156; Aerofloat 25, 0.018; pine oil, 0.088. Cyaniclation. The cyanide plant has a capacity of 300 tons per day. The flotation concentrates are reground in two stages, each stage employing a 5- by 16-ft. Allis-Chalmers tube mill loaded with 30,000 lb. of 17~-in. steel balls run in closed circuit with a 6- by 30-ft. DOlT duplex classifier. The primary-mill discharge at 52 per cent moisture is diluted with mill storage solution, overflows the classifier at 7.5 to 1 dilution, and after being mixed with secondary-mill classifier discharge at 1.4 to 1 dilution, passes to a 20-ft.diameter bowl classifier from which the reciprocating rakes have been removed and a bottom cone discharge installed. The underflow passes to the secondary mill. The overflow is agitated at 4.3 to 1 dilution in a series of TABLE

47.

SIZING ANALYSIS OF BOWL OVERFLOW

Microns

+ 56

-56 -40 -28 -20 -14

+ 40 + 28

+ 20 +

14

+ 10 -10

Per Cent

0.13 2.98 13.90 14.22 11.56 9.86 47.35 100.00

four 24- by 20-ft. Dorr agitators, where a portion of the bottom discharge is recirculated back through the bowl classifier. Following agitation in this series of agitators, the pulp passes into a second bowl classifier similar to the first, where the dilution is increased to give an overflow a 10 to 1 dilution and an underflow that passes back to the secondary mill for further treatment. The whole object of this circuit is to grind preferentially and contact the heavier and coarser portion of the ore with cyanide until practically all of the product passes 325 mesh. An infrasizing of the second bowl overflow is given in Table 47. This overflow goes to a 50- by 14-ft. Dorr tray thickener, from which point clear solution is withdrawn for clarification and precipitation. The thickened underflow is given three stages of filtration with about 20 hr. agitation in 24- by 20-ft. DOlT agitators between the first and second and second and third stages, respectively. Filtration is on six 8.5-ft.-diameter, eight-disk American filters. Water wash on the first filtration step has been found to assist extraction materially. Barren solution is used on the last two stages.

311


About 2500 tons solution per day is clarified and precipitated by the Merrill-Crewe system. Approximately 150 tons barren solution is discarded per day to balance the addition of wash water on the first-stage filters and also to avoid solution fouling. Reagent consumption; in pounds per ton are, is cyanide, 0.78; lime, 1.32; zinc dust, 0.084; lead acetate, 0.035: Solution strength is maintained at 3.0 lb. cyanide per ton and 1.60 ll~. per ton lime in the primary regrind mill, with pregnant solution at 1.80 lb. N aCN and 0.25 lb. CaO per ton of solution. Typical metallurgical results and percentages of recovery are given in Table 48. Pamour Porcupine Mines, Ltd. (Type IVc). This property has one of the most modern mills in the Porcupine district. Capacity is 1600 TABLE

48.

METALLURGICAL DATA AT McINTYRE PORCUPINE'"

Metallurgical results

Value, gold at $35 per oz.

Heads ........................ Flotation tails ................ Cyanide residue ............... Soluble loss per total residue .. Combined tails ................

$10.344 0.363 0.779 0.204t 0.436

Stage

Flotation: Unit cells Flotation circuit Total Cyanidation : Primary grinding Agitation and regrind Filters, etc. Total Total over-all recovery

Recovery 1946-19'17, per cent

75.00 21.90 96.90 65.00 31.65 2.20 98.85 95.79

* Representative

of results during the fiscal year ending June, 1947. t $0.024 per ton of mill feed.

tons per day. The ore is high-silica conglomerate carrying 2.0 per cent pyrite, 1.0 per cent pyrrhotite, and a low gold content (about 0.10 oz. per ton) associated principally with the sulphides. Primary crushing is done underground. A conventional two-stage surface crushing plant includes a standard Symons crushing to % in. size and a short-head Symons, closed-circuited with Niagara screens with a 2- by h6-in. opening, making a mill feed that is stored in two 750-ton ore bins.. The plant hanclles 120 tons per hr. The mill flow. sheet involves flotation to make a final tailing with the regrinding and cyanidation of the flotation concentrate, but it possesses the rather unusual feature of stage grinding with flotation steps between the stages. An over-all gold recovery of 91 to 93 per cent is made with this treatment scheme. Gold recovery in the flotation circuit is 94 to 95 per cent.

312


Primary grinding is done in two 9- by lO-ft. Dominion ball mills closedcircuited with 12- by 25-ft. DOlT classifiers with overflow at 40 pel' cent solids. Hydraulic traps are installed at the mill discharges to collect any TABLE

49.

CI-IE1IHCAL ANALYSIS OF PRODUCTS AT McINTYRE PORCUPINE

Mill feed Item

14-month 12-month average, average, 1936 1946-1947

8.21 Iron pyrite, per cent .............. Silica, per cent .................... 54.92 6.28 Ferric oxide, per cent ............. Aluminum oxide, per cent ......... 14.18 Calcium oxide, per cent .. " ....... 4.90 Magnosium oxide, per cent ........ 4.33 Sodium and potassium oxide, per 2.85 cent ............................ Gold dollars per ton at $35 per oz ... $13.05 2.84 Specific gravity (dry ore) .......... Lead sulphide, per cent .... , ...... · ... Nickel sulphide, per cent .......... · "Pyrrhotite, per cent ............... .... Chalcopyrite, per cent ............ '" . .... Sphalerite, per cent ........ , ...... Arsenopyrite, per cent ............ · ... TABLE

50.

6-month

I-month

1936

12-month average, 19·16-1947

6.81 52.76 6.43 14.32 5.25 4.69

0.10 58.28 6.79 15.07 5.61 2.82

0.15 58.83 6.11 14.19 5.14 4.65

56.67 18.80 3.53 8.07 3.24 2.50

2.15 $10.34

2.95 $0.406 2.72

2.25 $0.363

$89.00

... .

. ...

....

average,

average, 1947

Trace 0.05

. ...

0.34

0.51 0.37 0.37

GRINDING JYIILL DATA AT PAMOUR PORCUPINE

Item

Primary grind

Secondary grind

Size of mill, ft ........................... Connected hp ........ , .................. , Ball charge, tons ....................... , Size of ball, cast iron, in................. Mill speed, r.p.m......................... Mill discharge, per cent solids ........... Ball consumption, lb. per ton* ........... Liner consumption, lb. per ton .......... ,

9 by 10 350 34 3 18.5 80 1.8 0.14

5 by 14 125 9 1 28 70 0.79 0.06

* Based

Flotation concentrate

Flotation tailings

Concentrate regrind

7 by 12 175 11.5 1 22 62 0.4 0.02

on original feed of 1600 tons per 24 hr.

free gold. Following flotation in three banks of six No. 24 Denver cells, which make a final concentrate and a middling for cleaning, the tailing passes to a 32-ft.-diameter desliming thickener which produces a slime overflow and a sand underflow. The slime overflow averages 33 per cent


313

of the weight of the deslimer feed and is approximately 99 per cent minus 200 mesh. It is floated in one bank of six No. 24 Denver flotation cells, which make a rougher concentrate for cleaning and a tailing for discard. The sand discharge from the 32-H. desliming thickener is ground in two 5- by 14-ft. Dominion tube mills in closed circuit with a 13-ft. N orandatype elutriating classifier. Overflow from the classifier is floated in two TABLE

51.

SIZING ANALYSES OF GRINDING PLANT PRODUCTS AT PAMOUR PORCUPINE*

Primary circuit

Tyler mesh

Mill feed

3 4 8 14 20 28 35 48 65 100 200 -200

17.4 17.2 22.9 13.8 4.3 3.3 2.5 2.6 1.8 1.7 2.6 9.9

Secondary circuit

Mill Class discharge overflow

0.9 3.1 5.2 4.6 6.1 11.5 13.5 12.2 8.7 10.6 23.7

0.5 3.4 8.7 11.9 19.8 55.7

Classifier feed

8.7 9.6 11.6 19.0 51.1

Ball-mill Flotation discharge tailing

2.8 6.5 13.3 26.8 50.6

Flotation Size

+ 65 mesh + 100 mesh + 200 mesh - 200 mesh

concentrate, per cent

9.0 10.2 21.4 59.4

Size

o to 10 14 20 28

+

* These figures

to to to to 80

10 microns ....... 14 microns ....... 20 microns ....... 28 microns ....... 80 microns ....... microns ..........

0.2 1.7 7.4 23.3 67.4 Bowl overflow, per cent

60.3 17.5 13.6 5.9 2.6 0.1

are based on a tonnage rate of 1626 tons per day.

banks of six No. 24 Denver flotation cells, which make a rougher concentrate for cleaning and a tailing for discard. Low-grade rougher flotation concentrates are cleaned in one six-cell bank of No. 24 Denver cells with cleaner tailings returning to the primary grinding circuit and concentrates going to the concentrate grinding circuit. About 60 tons per 24 hr. of concentrate is reground in a 7- by 12-ft. Dominion ball mill closed-circuited with a 32-ft. N oranda elutriating classifier which overflows at 8 per cent solids. The concentrate is then dewatered in a 32-ft. thickener and passes to two 11- by 20-ft. Noranda-type agitators,

314


lime being added to the first, and cyanide to the second. Washing is accomplished in a five-stage C.C.D. system using 32-ft. bottom-drive N 0randa superimposed thickeners and 11- by 20-ft. agitators. The conventional system of clarification, precipitation, and cleanup is used. Hallnor Mines, Ltd. (Type Vc). Although this property adjoins that of Pamour Porcupine, the milling problem is somewhat different, and a very interesting flow sheet has been developed. About 350 tons per day of an ore of variable gold content (0.1 to 0.7 oz. per ton) is treated, and briefly the treatment scheme consists of flotation, followed by regrinding and cyaniding of the concentrates as at Pamour, but in this case the flotation tailings are first deslimed to eliminate about one-third of the total tonnage as a fine slime carrying negligible values and then cyanided in a circuit which receives the residue from the concentrate cyanidation section. Conventional two-stage crushing is used, delivering a minus %-in. feed to the 600-ton mill bin. Grinding is done in an -8- by 8-ft. ball mill closedTABLE

52.

REAGENTS USED AT HALLNOR

Reagent

Amyl xanthate Pine oil Soda ash. . .

Lb. per Ton

0.06 0.18 , 0.40

circuited with a 20- by 6-ft. rake classifier. An 18- by 24-in. Denver mineral jig is installed between the mill and classifier to remove coarse gold. The classifier overflow at 62 per cent minus 200 mesh is floated at 31 per cent solids in two banks of eight No. 24 Denver cells arranged in series, with middlings returning to the head of the cells and a final concentrate removed from the first three cells of each bank. The reagents used are given in Table 52. Treatment oj Tailings. The flotation tails pass to a 20-ft. desliming classifier, making an overflow at approximately all minus 200 mesh which is discarded and an underflow passing at 62 pel' cent solids to a series of three 14- by 30-ft. agitators, followed by three stages of countercurrent decantation in a 30-ft. N oranda superimposed thickener. Treatment oj Concentrate. The flotation concentrate is first thickened in a 23-ft. thickener and then passed to 10- by 30-ft. agitator, which is closedcircuited through the cone discharge with a 5- by 8-ft. regrind mill, while a second closed circuit from a discharge point higher up in the agitator is arranged through a 23-ft. superfine classifier, the sand discharge of which returns to the ball mill and the overflow at all minus 325 mesh passes to a 23-ft. thickener before the secondary stage of cyanidation. This consists


315

of a 14- by 30-ft. agitator and two stages of C.C.D. in a 23-ft. tray thic-kener. The underflow from this thickener passes to the third agitation stage of the tailings-treatment section, so that, in effect, the concentrate receives three successive treatment steps. Primary grinding and flotation are, of course, carried out in a fresh-water circuit, while cyanide is added at the head of the concentrate and tailingstreatment sections and regrinding is done in cyanide solution. Pregnant solution is taken off the first thickening stages in each section, a portion of tailings pregnant being advanced through the second stage of concentrate C.C.D. system. Clarification is done in 5- by 8-ft. leaf clarifiers" and precipitation in two 30-sock presses. Barren solution is returned to the third step of the tailings C.~.D. system. Black cyanide is used, the consumption, equivalent to 1.0 lb. per ton N aCN taking place almost entirely in the concentrate section. Lime consumption amounts to 4.0 lb. per ton, litharge 0.09 lb. per ton, and zinc 0.036 lb. per ton of ore treated. Lime and cyanide strengths are maintained at about 1.0 per ton of solution in each case, and an over-all recovery of 97.5 per cent of the gold is made. Hollinger Mill (Type Vc). This mill, for many years the largest in the Canadian north, has a maximum capacity of 4600 tons per day. A description of the operation was given by P. D. P. Hamilton in 1934 in Trans, 112, A.l.lII.E., 624, and in the Hollinger edition of C.lll.J., September, 1935, the various departments of the mine and mill were very completely described by different members of the Hollinger staff. Since 1935 the old crushing plant has been replaced by a new and very up-to-date installation-s-one of the finest in Canada-and a unique, circular ore bin of 10,000 tons' capacity has been constructed- (see Chap. III). Various other changes have been made in the milling flow sheet, most outstanding of which was the change-over from two-stage grinding using opencircuit rod mills followed by pebble mills in closed circuit to a single-stage closed-circuit grind. For this purpose the rod mills were converted to lowdischarge ball mills by placing specially designed screens at the discharge ends. The new crushing and grinding plants have been described by D. C. McLaren in C.lll.J., June, 1944. A comparison of the results obtained by the use of low and high pulp-level discharges from ball mills was given by the Hollinger mill staff in 1937 in Trans. 40, C.1I1.I., 85, 325, and also in Trans. 46, 1.111. and 111. Additional information was given by H. W. Hitzrot in the July, 1939, issue of T.P. 1088, A.I.1I1.E. Mining Tech. The ore is a siliceous schist of medium hardness with a specific gravity of 2.8. It contains 4 per cent pyrite and very minor amounts of other sulphides. For the past few years the mill heads have averaged slightly better than 0.235 oz. per ton. Silver is alloyed with the gold in the pro-

316


portion of 16 parts of silver to 84 parts of gold by weight. Grinding is done in cyanide solution, and the extraction of gold in the grinding circuit at the classifier overflow is 58 pel' cent of the head value. The following description has been brought up to date as of November, 1948, through the courtesy of the management. At this time the milling rate is 3700 tons pel' day. The ore is crushed to pass through a slotted screen with a clear opening of 0.205 by 0.50 in., with the slots set at right angles to the flow. The minus ;i-in. feed is fed to three or four ball mills, each 6;~ ft. in diameter by 14;~ ft. inside new liners. Each mill is served by a 12- by 27-ft. Dorr FX classifier, and the circuit is closed with a pump. The grinding unit has a capacity of 1200 tons pel' day with a circulating load ratio of 4: 1, grinding to 1 pel' cent plus 48 mesh, 60 per cent minus 200 mesh. The ball load is maintained by feeding 2;~-in. heat-treated forged-steel balls, hardened to 550 to 575 Brinell. In 1947 the consumption of ball steel was 1.13 lb. per ton milled. In brief, the general plan of treatment after crushing consists of grinding in cyanide solution, table concentration of sulphides, concentrate regrind with extra concentrate agitation, combined agitation of table tails and concentrates, and two-stage filtration for the major part of the tonnage, with three-step decantation followed by one-stage filtration for the remainder. Concentration and Concentrate Treatment. The overflow from the Dorr classifiers is pumped to 88 double-deck Deister slime-concentrating tables, where approximately 65 per cent of the pyrite and 80 per cent of the gold in the table feed are removed. About 525 tons concentrate is produced from 3700 tons feed, giving a concentration ratio of 7: 1. The concentrate is given a further grind to set free gold that is finely disseminated within the pyrite. This is done with one of the same low-discharge mills loaded with l-In. heat-treated forged-steel balls, in closed circuit with one 19-ft.diameter Dorr bowl classifier, overflowing at an over-all fineness between 80 to 85 per cent minus 325 mesh. Ball-steel consumption was 0.56 lb. per ton of original feed in 1947. The concentrates are thickened in three 40- by 15-ft. Dorr tray thickeners, the overflow going to the mill circulating solution and the underflow, at 56 per cent solids, going to four 20-ft.-diameter by 24-ft. Dorr agitators in series of two. Agitation time is 36 hr., after which the agitator discharge joins the thickened table tails for additional agitation. Lime is added as milk of lime in cyanide solution to the feed of all the ball mills, and the cyanide is added to the overflow from the bowl classifier in the concentrate grinding circuits. As a consequence, the solution in the concentrate section is higher in cyanide and lime than the main mill solution. The solution at the head of the concentrate agitators averages 1.2

317


lb. NaCN and 1.1 lb. CaO per ton, as against 0.75 lb. NaCN and 0.75 lb. CaO per ton for the pregnant solution. . Table-tails T;·eatment. The table tails are thickened in ten 40- by 15-ft. DOlT tray thickeners. Part of the overflow is sent to precipitation, along Cyanide solution

l

.

!

19,000-to+ mill bin

Milk oflime to ball mt'lls

",,;gbf"" bells

Four 6'-6/1x 14'-6/1low-discharge ball mills closedcircuited with four 12'x 27' QuadrUfz/ex Dorr classifiers

t

J

---,-

Vibrating screen (Wood chip. removal)

110 double-deck Deister concentratin tables

f

(liJll-ngs)

(Concentrate)

12 Primary 40'x 15' siJgle tray Dorr thickeners

Dorr bowl classtier closed circuited with 6'-6"x 14'-6" low-discharge ball mill

(Overflow)

3 Single tray-Dorr thickeners

/

'\.

•

:,

4 Dorr agitators 20 x24 in series of two

L

Clarification and _l2'ecl/2.ltation

l .

Meltlnnfurnace

_':7._

f

(Underflow)

-~-

t

(Underflow)

(Overflow)

~

.

~

6-Dorragitators2o'x24' ~ • 20-Pachuca agitators 15'x45'

r=

Pulp.. dls-tautor. -

Barren solution

Cyanide storage ona al(mill CirCUits

Barren solution

.

~

t

5 Rows of 3 stepcounter-current decantation traY.. thickeners

6-Primary" Oliver filters 14'xl6'

t-'1 Renulflers

to_

r - - -...-:::::.::-~SEec~o!!..!.rJ~do,:!!ry"Oliver ftYters

/4'x 16'

----=---t . Renulfler -1~

Water wash

Tailings to waste-

FIG. 73. Flow sheet of the Hollinger mill, Ontario, Canada.

with part of the filtrate from the primary filters, to maintain a precipitation ratio of 1.1 tons of solution to 1.0 ton of ore milled, and the remainder joins the circulating solution. The underflow at 54 per cent solids, along

318


"with the concentrate agitator discharge, is pumped to four rows of five 15-ft.-diameter by 45-ft. Pachucas. Agitation time is 20 hr. The Pachuca discharge is split two "ways. Up to 3400 tons of solids is sent daily to six primary 14- by 16-ft. Oliver filters, and the remaining tonnage is sent to one or more rows of three 40- by 15-ft. Dorr tray thickeners in C.C.D. The primary filter-cake discharge is repulped with barren solution, combined with the thickener underflow discharged from C.C.D., and filtered on eight secondary 14- by 16-ft. Oliver filters. Part of the filtrate from the primary filters goes to precipitation and part to mill circulating solution. .Part of the filtrate from the secondary filters is used as low-grade wash solution in the decantation thickeners, and the remainder joins the circulating solution. The cake on the primary filters is given a barren-solution wash, while that on the secondary filters is given a barren-solution wash followed by a water wash. The mill operates on a 6-day-week schedule and treated an average of 3627 tons daily during the 40 weeks ending Oct. 6, 1948. The loss of gold in the tailing amounted to 0.00677 troy oz. per ton of ore milled, of which 0.00042 oz. was the loss in dissolved gold. The recovery of gold was 97.1 pel' cent. Total milling costs, including underground crushing, refining, and tailing disposal, for the same period were 77.29 cents pel' ton milled, of which 37.90 cents per ton were labor costs. A cost-of-living bonus and social-service costs are not included. Cochenour Willans Gold Mines, Ltd. (Type IVr). This mill in the Red Lake district was started up in 1939 as an amalgamation and cyanide plant, flotation of the cyanide tailings being installed in 1940 to recover values locked in the sulphides. Concentrates were at first shipped to Tacoma, but this was later discontinued owing to difficulties in transportation, and they were stockpiled at the mine while the possibilities of local treatment were studied. In 1946 it was decided to convert the plant to an all-flotation operation with roasting and cyanidation of the concentrates. The flow sheet of the present mill, which is hanclling 220 tons pel' day, is shown in Fig. 74. The ore is composed of about 60 per cent silicified rhyolite, the remainder being a complex of cherty quartz and carbonates carrying free gold and gold-bearing sulphides. The gold content averages about 0.4 oz. per ton but in July, 1947, reached 0.67 oz. per ton. The ore is crushed to % in. and ground in a 6- by 8-ft. Allis-Chalmers ball mill, closed-circuited with a Dorr DSFH classifier, overflowing at about 70 per cent minus 200 mesh. A Denver duplex jig in the millclassifier circuit collects some coarse free gold, while the classifier overflow passes over two blanket tables 45 ft. long before thickening in a 24- by 12-ft. Dorr thickener prior to flotation.

319


Flotation is carried out at 10 to 1 dilution in a series of 14 Denver cells connected to give primary roughing, primary and secondary' cleaning, secondary roughing, and scavenging. The tailing is sent to waste. Concentrate is thickened and filtered,ancl the cake fed to the Dorrco FluoSolids Mill are bins

t

6 ' X8 ' Allis Chalmers ball mill

C

t

12'x18' Denverdu/2lex jjg_-----'l~

t

5'x 23~4 "Dorr D.5.F.1:I. classifier

t

2-45'long blanket tables in /2arallel ~

t· t

Pine 017 Xanthale CuS04 - - - ,

.

24' xl 2' Balanced Iray.Dorr Ihickener

I----l~~ 2- 6'x6' Denver condilioners

NaOH

t

.s:

14 Denver flolalion cells

(TJ7s) I r - - - - -........',

To

wasle~

,

II x 8 ,

I

'E

'

Denver thickener I

t

4 x 4 Denver filler

•

Dorr FluoSolids roaster

t

----.-

2-Blankel tables ------i~

NaCN

Cao I

)0

2"16'xI7' Dorr agjtators in series

t

24' x 12' Dorr Irax thickener

t

e'» 10' Oliver

f/lter

------r-Re/2u/;2er

To waste~--Ir----'--------'

3'x4'Grindingbarrel II

Vacuum leaf clarifier

t

r-

Merrill- Crowe /2recip-jlation PreClQilale shi!2/2ed

/I

f--

24 x 30 Amalgam barrel

f 'f--

Denver Mercurx se/2aralor Amalqam /2late

•

Amalgam ship-/2ed

FIG. 74. Flow sheet of the Cochenour Willans mill, Ontario, Canada.

320


roaster. A ratio of concentration of 27: 1 is obtained (see Chap. X for a description of this furnace and roasting operation). After repulping in a quenching tank, the calcine is passed over two more blanket tables and agitated at 36 per cent solids for 56 hr. in two 16- by 17-ft. DOlT agitators connected in series. The pulp is then thickened to 50 per cent solids and filtered on an 8-by lO-ft. Oliver filter using a barrensolution wash. The pregnant solution taken off the thickener is clarified in a leaf clarifier and precipitated in the conventional Merrill-Crewe system. TABLE

53.

METALLURGICAL DATA AT COCHENOUR WILLANS

Gold Recovered, * Per Cent

Item

Amalgam in mill circuit Calcine amalgam and cyanidation Total

, ,

52.3 40.3 92.6

Chemical reagent consumption in this plant: Lb. pel' Ton of Ore Treated July, 1947. 6290

Flotation

Ton Milled

Pine oil. Xanthate (Z-6) Sodium Hydroxide (NaOH) Copper Sulphate (CuSO.j)

0.14 0.27 0.54 0.37 Lb. pel' Ton of Ore Treated

Cyanida/ion

Cyanide. . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Lime Lead acetate , Zinc * First 6 months, 1948.

0.66 0.75 0.039 0.039

All jig and blanket concentrates are reground and amalgamated in an acid circuit, and the amalgam collected in a mercury separator and on an amalgam plate. Both amalgam and gold precipitates are shipped to refiners. Cyanide strength is maintained at about 1 lb. per ton of solution, and the alkalinity carried at pH 10.6. Some 500 tons of solution is precipitated each day. ZINC-GOLD ORES

Golden Manitou Mines, Ltd. (Type IVa). As pointed out by the authors" of an article on this property, the hundred-mile chain of producing mines, which reaches from the Kirkland Lake area in northern Ontario 3 Andrew Robertson and A. Livingston, "Golden Manitou-New Zinc Producer," Vol. 145, No.3, March, 1944.


321

on the west, across the provincial boundary to beyond Val d'Or and Bourlamaque in Quebec on the east and continues to be so productive of go1cl, is marked at its 60-mile point by the base- and precious-metal producers of the N orarida district and again at its far-flung terminus by the interesting zinc-geld-silver deposit of the Golden Manitou mine. . The ore, which is a mixture of pyrite and sphalerite with occasional streaks of galena and other minerals, carries (1944) 7.5 per cent zinc, 0.05 oz. go1cl, and 3 to 5 oz. silver. The 1000-ton mill makes a primary-flotation concentrate which is reground and cyanided. The cyanide residue then joins the primary-flotation tailings for flotation of" the zinc, the latter being brought up to grade in a special deleading cleaner circuit that follows the zinc circuit. Cyanidation is carried out in a conventional thickening and agitation circuit, followed by two-stage filtration. COPPER-GOLD ORES

Noranda Mines, Ltd.' (Type IVse). This 3000-ton mill treats a massive sulphide ore high in pyrrhotite. The ore averages 4 to 8 per cent chalcopyrite, 20 to 30 per cent pyrite, 50 to 60 per cent pyrrhotite, 0.12 to 0.20 oz. gold and 15 to 20 pel' cent insolubles, The flow sheet includes stage grinding and stage flotation of the chalcopyrite, which is smelted for recovery of copper and gold. Flotation is carried out in an alkaline pulp, which is aerated to counteract the reducing property of the pyrrhotite. It has been found that this aeration step is necessary after each stage of grinding and prior to copper flotation. The pyrite rejected from the tertiary copper circuit is floated and cleaned. It then passes to a rotreatment section where the remaining chalcopyrite and pyrrhotite are further reduced in a circuit which involves aeration, flotation, and regrinding before cyanidation- (see Chap. XIII for discussion of the methods involved). Cyanidation is carried out ill a series of six N oranda-type agitators followed by countercurrent decantation washing with pregnant-solution removal from the first-stage thickener for clarification and precipitation. The over-all recoveries in this plant by flotation and cyanidation are 96 to 96.5 per cent copper and about 85 per cent gold. UNITED STATES

The United States stands in fourth place ill world gold production and second place in world silver production, with a total of approximately 1,500,000 oz. gold and 20,000,000 oz. silver according to 1946 statistics. This is about one-third of the prewar production. Taggart, Handbook of Mineral Dressing, Sec. 2-65, Wiley, 1945. C'. G. McLachlan, H. L. Ames, and R. J. Morton, "Cyaniding at Notanda;" Trans. 49, C.I.M. and M., 91-122, 1946. 4

5

322


Gold mining? in the United States experienced more adverse effects from the United States participation in the Second World War than any other large mineral industry. Silver production was also seriously affected inasmuch as gold mining is a substantial source of by-product silver, although a large part of the output is derived from base-metal mining." The TABLE

54.

MINE PRODUCTION OF GOLD IN THE UNITED STATES IN

1946 BY

STATES

AND SOURCES, IN FINE OUNCES, IN TERMS OF RECOVERED METALS"t

State

Placers

Dry and siliceous ore

Copper are

Zinc-lead, zincLeadcopper, copper Zinc ore are and Lead zinc-leadare copper ores

Total

-Alabama ....... 1 Alaska ......... 220,708 6,003 ··0 ... 3,564 61,347 Arizona ........ 398 311 California ...... 269,772 79,883 131 Colorado ....... 20,172 106,007 21 .... - . .. ··0 . Georgia ........ Idaho .......... 20,123 17,027 60 Montana ....... 27,986 27,802 11,590 Nevada ......... 14,786 28,973 43,260 1,384 1,566 New Mexico .... 10 Oregon ......... 16,502 1,096 ..... . Pennsylvania ... 1,150§ . ..... South Dakota .. .... .. 312,247 ..... . Tennessee ...... ..... . I·· ..... . 95 1 .... .. Texas .......... o. Utah........... 7 19,075 133,454 165 Vermont ........ Washington ..... 101 18,686 3 Wyoming ....... 17 88 ...... •

•

0

0

••••

0

•••••

••••

•••

•

0

0

••••

0

•••••

TotaL ........ 590,604 621,836 253,132

...... 70 546 73 111 1,275 ...... 3,271 3 2,093 1,880 •

••

0

••

•

••••

••••

0

•

••

0

0

•

•••

0

••

0

.0

••

0.

•••

0.

•

1 107 634 ...... 371 179

•

••••

0

•

...... 12,985 2,312:1: 12,327 '0."'

40 68 103 743

•

5,617 2,427 3,187 127

•

.0

•••

•

.0

•••

'0' •••

0

••

• 0' • • •

0.0

•••

.0··0 .

"0 '0'

.0

0.0

•••

.0

•

••••

0

•••••

•

•••

••••

0

••

. ..... · ..... 7 1,300 312 ••••

••• '0'

0

••

0

0

. .....

•

•••••

0.0

•••

•

'0' •••

12 ...... 0

0

1 357

•

••

••.•

•••

6 •

•

0

•••

"0

•••

0

0

•

24,028 . ..... 32,360 . .....

- - --- --6,381

389

6,793

1 226,781 79,024 356,824 142,613 21 42,975 70,507 90,680 4,009 17,598 1,150 312,247 95 9 178,533 165 51,168 105

95,370 1,574,505

" Minerals Year Book, "Metals, Gold and Silver," by C. W. Merrill and H. M. Meyer, 1946. t Philippine Islands and Puerto Rico excluded. :I: Includes gold recovered from pyritic ore and tungsten ore. § From magnetite-pyrite-chalcopyrite ore.

mine production of gold in the United States increased 65 per cent' compared with 1945, but except for the war years, the 1943-1945 output was the smallest since 1885. Silver output declined 21 per cent in 1946 and, except for 1932, was at the lowest point since 1872. Of the 25 leading gold-producing mines, 9 were placers worked by B

Minerals Year Book, Part II, "Metals, Gold and Silver," by C. W. Me;rill, 1945. p. 557, 1946.

7 Ibid.,

323


connected bucket dredges, 8 gold mines, and 5 base-metal mines; 3 produced more than one type of ore. The 3 leading mines contributed onethird of the total gold produced in the United States in 1946, the 9 leading mines accounted for over one-half, and a list of 25, 64 per cent: Only 4 of the 25 leading silver-producing mines depended exclusively on silver ore; ores valuable chiefly for copper, lead, zinc, and gold accounted. TABLE

55.

JVIINE PRODUCTION OF SILVER IN THE UNITED STATES IN

1946 BY

STATES AND

SOURCES, IN FINE OUNCES, IN TERMS OF RECOVERED METALS*t

Zinc-lead State

Alaska ......... Arizona ........ California ...... Colorado ....... Idaho .......... Illinois ......... Missouri ........ Montana ....... Nevada ......... New Mexico .... New york ...... Oregon ......... Pennsylvania ... South Dakota .. Tennessee ...... Texas .......... Utah ........... Vermont ........ Washington ..... Wyoming .......

Placers

Dry and siliceous are

Leaq,. Copper are

1,141 ..... . 33,656 62 166,0441,764,558 19,821 77,708 16,736 3,698 1,290,723 36,286 6,422 2,719,762 1,086 .... .. .... ..

5,354 5,982 2 ..... .

3,229 . .....

...... ... ... .. .... ...... . .....

...... ......

.... .. .... ..

452,9302,082,321 242,709 289,701 29,824 100,318 ......

3,641 ..... .

86,901

. .....

.... ..

copper are

Zinc are

.... .. .... ..

.... .. . ..... . ..... . .....

. ..... ...... ..... . ..... .

...... . ..... . .....

. .....

791 ...... 813 . ..... 76,870 49,620 54,056 2,315,100 ..... .

..... .

Total

lead-copper ores

6,996 . ..... ..... . ...... 41,137 21,886 12,294 1,262,784 956,860 ...... 133,841 137,685t 39,911 77 117,607 751,849 264,279 18,154 39,614 3,441,787 35 ...... . ..... 2,267 67,734 1,667 . ..... . ..... 35,259 ..... . 16,750 680,526 196,518 ...... 17,339 498,402 3,537 ...... 140,674 63,645 .. .... . ..... . ...... 15,786

57 ...... 7,887§ . .....

18,016 18 41,300 385,7471,237,060 35,275 .... - . 21 103,369 349 13 10 3 . .....

Lead are

zinc-copper, and zinc-

......

.

.....

41,793 3,268,765 1,342,651 2,240,151 6,491,104 2,302 69,401 3,273,140 1,250,651 338,000 15,786 6,927 7,887 86,901 18,016 42,922 4,118,453 35,275 264,453 26

14,433 ...... 708 145,573 . . . . . . ...... ...... . ..... -TotaL ........ 78,250 5,601,812 5,589,678 1,704,382 533,674 9,315,404 22,914,604 191,404

* Miaierals Year Book, "Metals, Gold and Silver," by C. W. Merrill and H. M. Meyer, t Philippine Islands and Puerto Rico excluded. t

Includes silver recovered from pyritic ore and tungsten ore. § From magnetite-pyrite-chalcopyrite ore.

for most of the silver production. The 5 leading mines contributed 26 per cent of the total silver produced in the United States in 1946; the 10 leading mines, 39 per cent; and a list of 25,60 per cent. One of the anomalies of war economy was the emergence of a copper district, West Mountain (Bingham), as the leading gold producer in the United States, surpassing Lawrence County (Lead), South Dakota, in 1943, 194L1, and 1945. However, in 1946 California regained first place

324


as a gold producer from Utah, which sank to fourth place. Second and third places were occupied by South Dakota and Alaska, respectively, these and California supplying 57 per cent of the United States output. The three leaders depended very largely on straight gold mines, both lodes and placers. The four leading silver states-Idaho, Utah, Montana, and Arizonaproduced 57 per cent of the domestic mine silver. The recovery of gold was divided fairly evenly among placer methods, amalgamation-cyanidation mills, and smelting of crude ores and concentrates. Almost 82 per cent of the domestic silver output was recovered by the smelting of concentrates, and nearly all the remainder by the direct smelting of ore. Alaska." In all, 13 dredges and 16 draglines operated for some portion of the year 1945. Production was reported from 18 lode mines, and 172 placer properties were active. TABLE

56.

ALASKAN GOLD PRODUCTION,

1945

Item

Gold, oz.

Silver, oz.

Lode mines ..................... Floating dredges ................ Placer ..........................

10,409 34,885 22,823

2897 3933 3153

Total. ............ , ...........

68,117

9983

This is only about one-tenth of the gold production of 1941, but in 1946 the production had jumped to 226,781 oz. Since 1880, the year of the "gold rush," Alaska has produced over 25 million ounces of gold. The Yukon Consolidated, which is the largest operating company and owns eight dredges and a hydroelectric plant, handled over 3 million cubic yards of material in the 1946 season with an output of $1,400,000 in gold. The Alaska Juneau Gold Mining Co., which operated a mill of 12,000 tons' daily capacity comprising ball mills, tables, and a flotation plant, closed down in 1941 owing to increased costs. Homestake Mining Co. (Type IIss). This is the largest gold producer in the United States, with a monthly production of over 7 million dollars in gold bullion (1948). Discovered in the early years of the present century, the treatment scheme has been varied, and as many as five separate mills have at different times operated on ore from the Homestake mine. Today all the milling and sand-leaching operations are located at the town of Lead, while the slimes run by gravity to a central plant located at Deadwood, several miles below Lead. The Homestake operation is of special interest today in that it is the only large mill in the United 8

Ibid., Part II, "Metals, Gold and Silver," by C. W. Merrill, 1945.

325


Millore bins 8,OOO-ton 'Live load) 180 staml2s(~600 to)

-·i-

r-

10-1'diam. dewatering canes 65° slofJ.e

[

3-e'» 12r~'Y rod mill

~-5" 10'AlII,

Clark Todd amalgamators

Clark Todd amalgamators

.3-6'-0"x

rod mill]

~8~4"Dorrclassifiers 1-4~6"x 21'-J"DorrclaSSifiers

lutITf~----------2-6'x 26'-8 "x 12' bowl classifier

6'x 31'-8'xI6' bowl classifier

(Sands)

(Overflow)

~

t

t

t

t

(Overflow)

(Sands)

"

[2-5'X 14' Marcyball m/~/s

t.

2- 5 x 14 Marcy ball m/~/sJ

2-6'x26'c8" Dorr classifier

6'x 23 '-4" Dorr classifier

Clark 70dd amalgamators

, :t OF. 6 -1 dewatermg cones

2-6'x33'-4'xCO"diam. bawl classifier

--.......,~

t

.

t

(Slime)

4~6'x2/~4 "Dorr classifier OF.

t

<, (Sand) f SandI2lantNo.3

8-44'diam. x 12' deep tanks.1I0-ton per charge

2-5'x14'A.C. tubemills

•

/ - 6 X21 '-8" Dorr classifier [

SandRlantNo.1 21-44 xII '-6 " tanks. 685-tons per charge

, , t 2-5 x 14 A.G. tubemills 8-Settling cones10'0D. x500slope (UnJerflowJ

(OVe'flowJl

I_

.

fI

~ Dorr thickener

70unit A 10-Settling cones 8' QD. x 50° slope

5- 9 '-6" mechanicalcones 9 ~6 "diom;

i

Slime {}lant 31-Merrillpresses

Tosandplants

FIG. 75. Flow sheet of the Homestake mill, South Dakota.

326


States to employ stamps and separate treatment of sand and slimes. This method has been retained because of the high metallurgical efficiency it gives on an ore that is somewhat difficult to treat. The presence of pyrrhotite and other minerals carrying ferrous iron demands special chemical and mechanical controls (see Chap. XII), and much thought has been given to this problem by the metallurgical staff. A new flow sheet involving all-slime treatment and the elimination of the stamps is now, however, under consideration. The present mill flow sheet is shown in Fig. 75. The 180 stamps take minus 27~-in. feed and are equipped with ~-in. battery screens. The

FIG. 76. Cyanide plant, Homestakc Mining Co., Lead, S.D. continuously since 1887.

This mine has operated

stamp-mill discharge after dewatering in 7-ft-diameter cones is ground in Marcy and Allis-Chalmers rod mills closed-circuited with Dorr classifiers with Clark Todd amalgamators in the mill circuit (see Chap. XI). Following this primary grind, the pulp is deslimed and reground in three separate units which have been installed at different periods and contain different equipment. Smaller amalgamation units are also installed in the ball-mill and tube-mill circuits. Unit A, with a rated capacity of 900 tons per day, deslimes first in bowl classifiers, regrinds the sand in two Marcy ball-mill-Don- classifier circuits, and makes a final desliming in two more bowl classifiers. Unit B, with a rated capacity of 1900 tons per day, follows the same general scheme as unit A but makes the final sand-slime


327

separation in the cone section instead of bowl classifiers. Unit e (1000 tons per day capacity) uses a combination of dewatering cones and Dorr classifiers with regrinding in Allis-Chalmers tube mills, the final separation being made in the cone section. The latter involves three stages of cone classification, with mechanical-hydraulic cones in the last stage. The sands are transferred to one of two sand plants, having a totai of twenty-nine 44-ft.-diameter by ll-ft. 6-in. tanks while the slimes are thickened in Dorr thickeners and flow by gravity to the slime plant which houses thirty-one 90-frame Merrill presses. This is the only installation of its kind in the world, in which both aeration and leaching are carried out in presses. oJ Details as to sand and slime treatment at the Homestake will be found in Chaps. VII and VIII, respectively, while a discussion of the cyanicide problem is given in Chap. XIII. At the time of writing, 52 per cent of the mill feed is treated in the sand-leaching section and 48 per cent in the slime plant. With a mill head of 0.4 oz. pel' ton gold, approximately 70 per cent is recovered by amalgamation. The sands assay about 0.19 oz. per ton, and the slime 0.10 oz. per ton gold. The total extraction is 96 pel' cent, of which 64 per cent is made by amalgamation. The reagent consumption per ton of ore treated is mercury, 0.065 Troy oz.; cyanide, 0.65 lb.; zinc dust, 0.06 lb.; and lime 2.20 lb. Cripple Creek Ore. Golden Cycle (Type IIss). At Colorado Springs the custom plant of the Golden Cycle Corporation has operated since 1907. The gold ores that are treated by roasting and cyanidation are the sulphotellurides received almost exclusively from Cripple Creek, according to L. S. Harner in I.e. 6739, UB.B. of M., 1933. Two types of ore come from the district, siliceous and basic, containing 76 to 87 per cent silica, 37~ to 4 per cent iron, 1.8 to 2.3 per cent sulphur, 1.6 to 5.2 per cent lime, and 0.5 to 1.4 per cent magnesia. The siliceous or oxidized ores are easily treated, but the denser and harder ores from depth give more difficulty. Cripple Creek mine ore averaging 0.40 to 0.80 oz. pel' ton gold and dump ore averaging 0.10 to 0.15 oz. are treated in varying proportions according to the flow sheet in Fig. 77. . The plant has had an interesting history. Up to 1929-1930 the ore treatment consisted of straight roasting followed by cyanidation, at which time eight or nine 70-spincUe Edwards roasters were in operation treating about 1000 tons per day. It was then decided to treat base-metal gold ores in addition to the regular Cripple Creek gold ores, and a flotation plant was added, the concentrates being shipped and the tailings filtered and cyanided along with the roasted gold ores. When the pi-ice of gold was increased in 1933, the treatment of base-metal ore was discontinued and

328


scheme of grading ores was adopted, with the higher grade ore going to the regular roasting plant and the lower grade floated in water, followed by roasting of the concentrates and filtering and cyaniding of the flotation tailings. At a later stage all the ore was floated. The next scheme was the development of flotation in a cyanide circuit using special activators, and during this period approximately one-third of the material roasted was flotation concentrate. During the Second World War lead-zinc-copper ores from the district were floated, the concentrates shipped, and the tailings cyanided. At the end of the war period, gold ores exclusively, were treated using again the technique of floating in cyanide solution, but at the present time with operations on a reduced tonnage basis and while plans for a new mill are being worked on to replace the present plant, the original scheme of direct roasting all the ore has been temporarily adopted. The plant has a rated capacity of 1500 tons per day, but at the present time not more than about 500 tons per day is being treated. Mter crushing to about 2 in., the ore, which is received in carload lots, is passed through a sample mill (see Fig. 13, Chap. IV). The rejects from this operation are then crushed in short-head Symons crushers closedcircuited with Hummer screens to approximately minus 3 mesh. The use of flywheels on the direct-connected drives to these machines is said to improve the operation in that it enables the machines to take momentary overloads more easily. The roasting operation is discussed in Chap. X. Grinding is carried out in cyanide solution in 6- by 6-ft. ball mills that have been successfully converted to rod mills, closed-circuited with an Akins classifier. The coarse (18 mesh) grind desired is difficult to classify, but the addition of slime to the classifier has improved the separation by increasing the pulp density, The mill discharge passes over Canton flannel blankets, and the concentrates are treated in amalgamation pans, the tailings from which are returned to the mill circuit. A sand-slime separation is then made in both Akins and Dorr classifiers, the sand analyzing about 5 per cent plus 20 mesh, 85 per cent plus 60 mesh, and the slime 90 to 95 per cent minus 200 mesh. The slime is thickened, agitated for 60 to 70 hr. in two stages with a thickening step in between, and then filtered in a 166-leaf (67~- by 9~-ft.) Butters plant. The sands are leached in 50- by 15-ft. vats, which are full of solution during filling with Butters and Mein distributors, while air is passed up through the filter bottoms. This promotes both aeration and displacement of slime, which passes to the thickeners of the slime circuit. A 7-day leaching cycle is used. The flow sheet in Fig. 77 clearly shows the solution flows and also illustrates the system of clarification

329


Miscellaneous dum/2 andmine ores 4'x it)' Hummer screens Qla''x Q9 "o/2eninfl. (Unirsize)

.

(Ove!size) 3-5Yz' short head Symons crushers

Stora(7/bins

-"r

a-Edwards drlex roasters --------~ _OIi_O~ confxor

rr2-6X6

Coloiron works'ad mNls

t

Ii

60"Akins classifier

I' I

(Rake) (Overflow)

t

f

II

Gases to spray washer and{Jottrell preeip/~ators solids",,""'" 10roosIe,

~

1

I

Canton flannel blankettables II x20 - k ..

I I

,.

i

I

t

I

cona

2-36"Akins and 2 Oorr classifiers f' (5;'-) (C' d.J

to amalgamator

t iI pans, Amalgam to I une »on / retort I , t i t , L..--50 Oorr trax thickener I,,---+- 50 x 15 leaching vats +---j II

III (IIF!'ma/ sounon ~ Arst , SOlutIon f .--:ld »on L.. f 1-, : r--'" 50x 15' transfer vats

t

-r-a~itator r--

'J J 41'x 26'0orr agjia,Or

I

...1

37'x25'Oorr

II ,I I

, .

f

.

J?

r-r--

Finalsolution I

I I

L

I

...I

I I

I

1

,.

-.--t-

r:

-r--.--

I

-,-

6-Merrill {JreciRilation jJresses

Sollosl

't'

Waste

,--

--------~

..J

Crowe vacuum sJ!§.tem

I

I I I

I I

. I

Gold solution tank

t"

t First solution

.:

I

1

2-Butters filters (Intermittent) I

t

I I I

t --t---~I .L ----

I I I II I I

30 'x 29' Oorr a~itator

-I

SolutIon: Overflowsolulton. I I and slirpe

I:

I

,

t . 'I

/

......5-30 Oorr trax thlckener---~ ISand . 35" x 20 Oorr a~tlator :r -f 30'x29 'Dorr agjTator

'

Solution

Waste L..

1--

I I

I

...1

Precipitates I

--.

r t

Rockwell furnace

~

Gra/2hite crucible Mint

FIG. 77. Flow sheet of the Golden Cycle Corporation mill at Colorado Springs, Colo.

330


used, 'which involves passing all pregnant solution from the slime plant through the sand plant before precipitation. The pH of the pulp is maintained at 10.6 to 11.0, and cyanide strength at 0.5 lb. per tone of solution. Most of the cyanide make-up in the form of 91 per cent N aCN is added at the zinc-dust feeder ahead of the presses, because rather high free cyanide is required for good precipitation. The ratio of precipitation is about 4: 1, and the resulting bullion is over 900 fine owing to the low silver content of the ore (10: 1 Au: Ag). Some 25 per cent of the gold is recovered by blankets and amalgamation and about 70 per cent by cyanidation. The reagent consumption in pounds per ton of ore treated is cyanide, 0.75; lime, 2.5; hydrochloric acid for filter leaves, 0.17; lead nitrate, 0.012; zinc dust, 0.145. Getchell Mine, Inc." This mine, a comparatively recent gold discovery, is in the old Potosi mining district, Humbolt County, Nevada. The ore body occurs in a replacement tuff shale bed following a basin range fault on the east slopes of the Osgood Range. The ore bed is highly siliceous with considerable carbonaceous material along with a clay gangue having the properties of bentonite. At one time two types of ore were mined, oxide and sulphide. The former was amenable to direct cyanidation, and 600 tons a day was handled in a sand-slime plant by more or less conventional methods. The sulphide ore, on the other hand, is more refractory, since all the gold is of micron size, and some of it is locked in fine sulphides. The ore contains from 1.5 to 2.0 per cent arsenic, which is present as orpiment and realgar with minor amounts of arsenopyrite. The orpiment and realgar are of later mineralization and are barren of gold. In the original plant 400 tons per day of this sulphide ore was roasted in a 77~-ft.-diameter by 260-ft.-Iong rotary kiln, ground in hot water to 65 mesh, and cyanided. Roasting eliminated 90 per cent of the arsenic and 70 per cent of the sulphur, and the gold extraction depended largely on the completeness of oxidation of the sulphide minerals. Much difficulty was encountered in thickening the slime fraction of the ore because the pulp tended to gelatinize during compression and low underflow dilutions were difficult to obtain without "island" formation when using a conventional mechanism. The original treatment scheme has been discontinued, and at present (August, 1948) the company is engaged in a construction program to change over and enlarge the old circuit to a flow scheme handling 1500 tons per day of sulphide ore. According to R. A. Hardy, consultingengineer.!? the present status and proposed flow sheet of ore treatment involve, 9 Fred Wise and C. W. Wark, "Metallurgy and Milling Practice at Getchell Mine," Trans. 153, A.I.M. and M.E., 1943. ' 19 Personal communication to the authors.


331

first, removing the non-gold-bearing, arsenic sulphide minerals, realgar and orpiment, by flotation. The arsenic flotation tails are classified into a sand and slime product at approximately 200 mesh. The slimes are subjected to a pyritic float. These flotation concentrates are returned to the sand fraction, and both are given a calcining roast. The- calcines are. cyanided by conventional methods of countercurrent decantation and zinc precipitation. The slime flotation tailings are cyanided and the gold values are recovered, after substantially total dissolution, by employing activated. carbon in counterflow to the pulps. Gold Mining in California. The "California Gold" issue of E. ancl .M.J., November, 1934, contains a comprehensive survey of the history of mining and ore treatment in this once-great gold-producing area of the world. During the 100 years following the discovery of gold in 1848, the total output of the state has amounted to about 2 billion dollars (at $20.67 per ounce). Placer and hydraulic mining accounted for most of the production in the early days, with drift mining and dredging assuming greater importance after the turn of the century. In 1910, a rather typical year, with the total production close to 20 million dollars, 72 operating dredges accounted for about 40 per cent of the output, with roughly 10 pel' cent from placer operations and the remainder from lode mining. The California gold-mining industry was, however, seriously crippled by the high prevailing costs, labor shortages, and other causes during and following the First and Second World Wars, so that today (1946) the gold production is only slightly more than 12 million dollars (at $35 pel' ounce) and only a few mills of any size are to be found operating in the whole state. A great variety of treatment schemes have nevertheless been used in the many well-known mills at one time in operation, and as a matter of historical interest the more typical of these are briefly summarized in Table 57 arranged according to districts. The gold ores of California range in kind from alluvial material carrying free gold to vein filling in which there are found both coarse and fine free gold and gold associated with a small amount of sulphides. In the socalled "Mother Lode" vein system, which stretches along a distance of 100 miles in the foothill country of five counties, the country rock consists of greenstone and slate, with variable amounts of carbonaceous schist. The gold occurs in quartz veins and in the enclosing slates. The principal sulphide is pyrite, with minor amounts of pyrrholite, arsenopyrite, galena, and others. As much as two-thirds of the gold can frequently be recovered by amalgamation, and it is usually unnecessary to grind finer than 80 to 100 mesh for satisfactory recovery. Tailings from the stamp amalgamation concentrator mills of the Mother Lode of California have been undergoing cyanide treatment for many

TABLE

57.

TYPICAL CALIFORNIA MILLING PRACTICE,

1936*

w

W

Property

Area

Tons per day

l--:l

Treatment scheme

Siskiyou, Shasta, Lassen counties

Mountain Copper Solomon Mines

700 Dry crushing; screening; sand-slime leaching. 300 Pebble mills; plate amalgamation.

Sierra and Nevada counties

Sixteen-to-One Mill Spanish

100 90

Zeibright Grass Valley and Nevada counties

Mother Lode

Idaho-Maryland

400

Empire-Star

300

Lava Cap

300

Beebe

400

Montezuma-Apex

250

Argonaut

250

Kennedy Carson Hill

Columbus Mono County (Bodie) Rand mining district

1000

Standard Consolidated King Solomon Yellow Aster

150 700

35 200 50 250

Ball mills; screens; riffles; tables; amalgamation. Ball mills and classifier; cyanidation ; thickeners; filters; zinc dust precipitation. Grinding, jigs, amalgamation, regrinding, and cyanidation.

a

Ball mills; riffles; tables; amalgamation; flotation; cyanidation of flotation concentrates; thickeners; filters. Stamps, jigs, amalgamation, flotation and amalgamation of flotation concentrates. Ball mills, unit cell; jig; flotation; concentrates to smelter; jig on tailings launder.

o

Hardinge-Hadsel mills and Dorr classifier; flotation; concentrates reground and cyanided; zinc dust precipitation. Ball mills; hydraulic traps in classifier circuit; corduroy; flotation; concentrates to smelter. Stamps; amalgamation in box and plates; Vanner tables; tube mills; shaking amalgamation plates; shaking tables; tailings to cyanidati on; concentrates to smelter. Stamps; amalgamation; tube mills; flotation. Stamps; amalgamation; hydrauli c traps; tables; concentrates to amalgamati on plates; all tailings to tube mills and tables with middlings reground; sand-slime separation and cyanidation; concentrates barrel amalgamated. Stamps; tables; flotation.

Z a

(Dump ore) Grinding, flotation and shipment of concentrate. Stamps; amalgamation plates; ball mill; flotation. Stamps; plate amalgamation; flotation.

* In general the mining areas covered progress from the northern part of the state through the central area to the southern part of the state. Dry-crushing equipment, usually various types of jaw crushers, are included in most instances ahead of the wet-grinding units.

~

Z

>-< tj

~

>-<

Z

~tj a o

t=J Z

~

~

o

Z

o

"j

o

~

t=J [J1


333

years. This material comprises both accumulations from earlier operations and tailings currently produced. Interest in tailing treatment was revived by the lOOO-ton plant erected by the Kennedy Company to handle a dump of 1H million tons. : The customer plant of the" Amador Metals Reduction Company received 250 tons tailings a day from the Argonaut mill, and the Central Tailings Works received 125 tons fromthe Central Eureka mill. According to M. N. Colman in E. and lJ1.J., November, 1933, a typical form of procedure was to separate all sand from slime by means of mechanical classifiers, leach the sand, agitate the thickened slime and filter it, and precipitate the gold from the cyanide soluJ tions on zinc dust. Carbonaceous matter, generally associated with the slime, was troublesome. A coal-tar product was mixed with the slime and prevented premature precipitation. The investigation of E. S. Leaver and J. A. Woolf published in T.P. 481, U.S.B. of M., 1930, will be remembered in this connection. They concluded that the gold lost in the mill tailings was largely in the remaining sulphides and carbonaceous material. The sand usually gave satisfactory recovery by direct cyanidation, but Leaver and Woolf preferred to float the mineral from the slime and roast and cyanide the concentrate. Evidently, some operators followed this advice in California, for flotation of old tailings was at one time extensively practiced. Comstock Lode. TIllS once-rich mining area neal' Virginia City in southwestern Nevada is today of historic interest only. The great fissure vein about 4 miles long was discovered in 1858 and up to the turn of the century had yielded over 300 million dollars in gold and silver. The highgrade ores occurred in great lodes, 01' "bonanzas," some of which carried several thousand dollars to the ton. The ores in the Virginia City area were essentially silver ores carrying as little as 1 :40 ratio of gold to silver, but the Gold Hill and Silver City ores yielded higher gold ratios, and some were to be regarded as gold ores. Following the period of primary production, a number of plants were actively engaged in the retreatment of lower grade ores and tailings from the earlier operations. By 1935, however, there were only a dozen small mills operating in the Comstock area, and at the end of 1946 only one enterprise, that of the Consolidated Chollar Gould & Savage Mining Co., was active. The milling operations, which were suspended in December, 1947, were started up again recently on a 400-ton-a-day basis. The original Comstock practice involved stamp mills followed by amalgamation on plates and by the Washoe pan-amalgamation process, which was developed in the district. Chemicals were generally added with the mercury, usually copper sulphate and salt and sometimes sulphuric acid and metallic iron. With the advent of cyanidation many retreatment

334


plants were erected, but in general, owing to the soluble salts present in the tailings, results were not too satisfactory. Later, flotation was incorporated into the flow sheets, but the same soluble salts tended to give variable and unfavorable results. Commenting' on the great variety of treatment schemes being used in 1935, Gardner and Carpenter" remark that the use of flotation alone on Comstock ores had not, to date, justified the faith and hope placed in it and that often amalgamation alone and usually cyanidation alone were its proved economic superiors, the latter giving at times nearly complete extraction but at a greater installation and operating cost. Where, however, the scale of operation justified a more complex plant, combinations of amalgamation, table concentration, flotation, and cyanidation might be superior to cyanidation alone. Knob Hill (Type VS).12 The ore now being handled in the 500-ton mill at Republic, Wash., is hard and abrasive. Ore minerals in the vein system are unusually fine grained, and free gold is seldom seen even in the richest concentrations. They occur most characteristically in concentric banding of cherty quartz, calcite, and adularia. Individual minerals observed include quartz, chalcedony, adularia, sericite, calcite, graphite, barite, gold, pyrite, realgar, marcasite, tetrahedrite, polybasite, pyrargyrite, and stibnite. The flow sheet is shown in Fig. 78. Monthly production amounts to about 4000 tons of concentrates, which are shipped to the Tacoma smelter. Milling follows modern gold-recovery practice, with removal of sulphide minerals by flotation and retreatment of tailings in a cyanide unit. Coarse ore from the 100-ton storage bin is reduced to the proper size for ball-mill feed (minus % in.) by crushing in two stages. Grinding to 5 per cent plus 100 mesh is done in a ball-mill classifier unit consisting of an 8H- by 9-ft. Traylor ball mill and an 8- by 27-ft DOlT classifier. Between these two units is a 42-in. Dorrco Pan-American placer jig. The finished classifier product overflowing at 15 per cent solids is treated in a bulk roughing and cleaning flotation step as is shown in the flow sheet. Finished concentrates, after being thickened and filtered, are stored in a 100-ton bin for shipment. Tailings go to cyanidation. The hutch product of the jig is treated in a 12-in. Pan-American pulsator jig, the hutch product of which goes to an amalgamation barrel. Bullion is shipped to the United States mint in San Francisco. Reagent consumption per ton of ore treated is copper sulphate, Aero11 "Present Day Milling Plants on the Comstock Lode," presented at the regional meeting, A.I.M.E., San Francisco, Oct. 3-4, 1935. 12 Abstracted from E. and MJ., Vol. 149, No.3, March, 1948.


335

Mine ore

"T

30 n x 36 Tray~awcrusher A.C. low head vibratin9. screen (Undetsize) ,

I

%"op"ening.

(Ovefsize)

"I. 2"28 TraY.!.QLlYfatofies

650-ton fteorebin

~ 8~/x 9' Trafor ball mill Q "' J ,

,

Pa~- American 42 "placer ry'i-p.~.

~

S8 (Hutch)

,

I

Pan-American 12" Jj'q.

f

(Hutch)

f

(Overflow)

(Overflow)

, I .. x27 Dorr class/fler

6-Cell fagkgren flot. machines I (Conc.) (To{!s) (Middlings)--

-r-

.

~l

,Amt/9.am barrel _To_1Z_ing.

'1'

,

6-Cell Denver sub-A machinesJ'

' t{TaLils~)_.:.:(M.:.:;1.::i:Jd.:.:.~/::.:n~gs~) . . . . _ _-E__~.,I

II' , • t B.Ullion ,0 mm

(Conc.)

t

Wemco thickener

-

--r-

6ft. a-disk Americanfl7ter 100-ton co!centrate bin

r

t

Concentrates to smelter

5 - Trax. Dorr

No. I underflOW]

3-26' t;tot""

~hickener- No. I Overflowto mill solution

~NO.2

!

No.5 Underflow To tOilinis dam

o

Overflow

Cto"J"'thn precJtafion {Barrenr:::::a-n-)----e,-U--Iton 10' To No.5 Iroy mini

Oneslage of Ihickening followedby four stages ofC.

c.o.

FIG. 78. Flow sheet of the Knob Hill mill at Republic, Wash.

336


float No. 25, and A.C.C. No. 404, each added to the ball mill, 0.20 lb., 0.10 lb., and 0.04 lb., respectively; DuPont frother B-23, to cell 1, 0.025 lb.; Z-6, to cells 1 and 4, 0.025 lb.; and sulphuric acid, to cell 1, 0.25 lb. The last-named reagent is added to maintain a pH of about 6.5. As to the cyanide plant, 1.5 lb. of sodium cyanide is added to agitator 1 and 8 lb. of lime to aerator 1. Zinc-dust consumption per ton of solution is 0.025 lb. Steel consumption is as follows: crusher parts, 0.05 lb.; cast-steel balls, 7 lb.: and ball-mill liners, 0.60 lb. The latter are of the Britannia type. Idaho-Maryland Mines Corp. The concentrate treatment section of this plant is typical of a small-scale operation handling less than 20 tons of mixed flotation and gravity concentrates by batch-cyanidation treatment (see Fig. 79). The flotation concentrates, which carry a large proportion of talcy gangue mineral, result from the treatment of about 1000 tons per day of ore from the Grass Valley district in California. Coarse gold is recovered by gravity and amalgamation methods prior to flotation. An interesting feature of the concentrate-treatment plant is the use of agitators in the grinding circuit to assist the dissolution of gold as it is released in an operation which grinds all the ore to about 2 per cent plus 325 mesh. The classifier overflow is thickened, with pregnant solution passing to a storage tanh: and the solids pumped to one of a series of seven batch agitators, where a treatment period of 1 to 10 days is provided depending on the results of periodic residue assays. The ore is given several changes of cyanide solution followed by decantation to the weak-solution, or stock, tame, the pulp then passing to a Dorr thickener and filter for dewatering and displacement of gold solution. The pregnant solution from the storage tame is precipitated in a Merrill-Crowe unit, and the precipitate melted down to bullion.

SECTION 2. CENTRAL AND SOUTH AMERICA CENTRAL AMERICA NICARAGUA

Gold mining has become the basis of this country's economy, and a number of important producers, described in some detail below, recover gold from a siliceous quartz ore which usually carries appreciable basemetal values. Cyanidation is the basic process employed. Transportation is one of the major problems in Nicaragua. The mines in the Atlantic Coast area depend almost entirely on air freight, but this makes them accessible at all seasons of the year. The mines on the western coast, however, which are served by a combination of railroad and highway facilities, are subject to seasonal


337

transportation difficulties, as the roads become almost impassable during the rainy souson.P

.

The following are the more important gold producers of Nicaragua, Companhia Minera La India (Type IIa). This is a 3pO-ton mill operation on a quartz reef in decomposed andesite of rather high clay' content. There are practically no sulphides and very little coarse gold in the ore. The milling circuit includes two-stage grinding with classification of the primary-mill discharge for elimination of clay slime from the TABLE

58.

MILLING DA'l'A: NICARAGUAN .,GOLD MINES

Item

LuLuz

La India

Neptune

Tons milled per day. . . .. .. . Head value, oz. per ton . Approximate gold recovery, per cent . Mill feed screen size; in . Primary classifier overflow minus 200 mesh.

1350' 0.127 Au 90 -0.25 to -0.375

330 0.30 Au 0.60 Ag 94 -0.375

650 to 700 0.28 Au 0.30 Ag 93.4 -0.625

40t

19.7

35.4

per cent

.

Secondary classifier overflow minus 200 mesh, per cent . Agitation time, hr . Dilution in agitators, per cent solids . Number of stages of washing . Cyanide solution strength, lb. per ton.: . . Cyanide consumption, lb. per ton per ore .. Lime solution strength, lb. per ton . Lime consumption, lb. per ton per are . Solution precipitated, tons per day . Solution loss, oz. per ton solution . Barren solution wasted tons per day . Costs, dollars per ton milled ' .

85 to 90t 50 to 60 6 and 17§ 3 C.C.D. 1.0 0.75 to 1.0 0.2 0.4 1100 0.003 500 1.075

70.0 to 79.1 30 to 40 35 to 40 3 C.C.D. 0.7 KCN 1.0 0.5 CaO 10.0 1250 0.003 1.50

62 to 76 40 50 (?) 3 C.C.D. 1.0 ('I) 3.02 0.6 4.5 2200

1. 75

* Feed to flotation. 40 to 50 tons per day of flotation concentrates cyauided.

t Flotation feed. t Reground flotation concentrates. § The first agitator receives the classifier overflow direct. ened feed.

The second and third agitators receive a thick-

sands that pass to the secondary mills. TIns primary slime, which is intensely flocculated in the high-lime circuit, interferes with effective classification and is by-passed to the cyanide circuit. In order to increase the mill capacity to 400 tons per day, consideration is being given to improving this primary classification and treating the slime in a separate circuit. The mill flow sheet is otherwise that of a straight C.C.D. cyanide plant, without filtration and employing the usual clarifying and precipitating equipment. Neptune Gold Mines, Ltd. (Type IIa) is milling about 675 tons per day of ore consisting of quartz stringers in decomposed (clayey) andesite 13 Abstracted from an article by Fern Kettel, E. and M, J., Vol. 143, No.8, August, 1942.

Ci:) Ci:)

00

o

BRUNSWICK MILL CONCTS.

CUSTOM CONCENTRATES

In20 cu.ft. minecars hauledby truck1.6miles

IDAHO MILL CONCTS. In 2 Ianboxes handtrucked100ft.

I

Flat.conct. IQ9 tons 3.38 azs. Jig conct. 3.8 itons3.30 ozs. -

Rat. cancts. 2.8 tans5.40 azs• •jig canCts. 0.5 tans11.69azs. Amal. Bbl.TIs. Qltons 1.38ozs. I

14.7 Ions 3.36ozs.

3.4 Ions 7.38ozs.

~ H

t:J

P->-3

H

o

Z

tJ> Z

t:J

o o

Z

o

PLATFORM SCALES Pipesampled manualfeed

t=J Z

~

~

o

/8.1 Ions4.12ozs.

Z

I

r- Stocksolulion

o

6 hours/day013 Ions/hour

";j

I

Pump

I 60% solids

..

I

'7~"""' .. '1 4 '-6 "x28 'xl t' D.S.C. B. ~~

Dorr bowlclossiYier

o

!;d

t=J

(j).

I/O GPM0130 % solids- 63.11ons 1.8lozs. Solulion heldat 3.0 tos.NaCn, 1.81bs CoO

AII agitalars run 24 hoursQerday" 19 Ions/lank 0130 %solids of 3.6 Sp.Gr.

~ ~

tel

12 Ions per lank 0130 %solids of 3.6 sf;. Gr.

~

tel

Z

o'"" TolalconsumQlion/lon concl. Cyanide Ibs: 12.15 Lime Ibs. 18.85 Sleel balls Ibs 5.63 Tolalcosl/lon. concl. Power and water I. 16 Supplies, mainl. Irspln.etc. 3. 25 Labor 3.56 Tolal Coslplus residues

7.97 /111.47

!

_n

-

--I ~ ---

';I

Flux

CO)

o

%wi. ofdrx..12f2!,.

Bullion Fineness

Borax g!ass Soda ash ~ Povd. sl'lica Mr!JJgQnese dioxide Sodium nilrole _FluorsQar

40

25 20 5 10 2

8 o

::0

tel

[fl_

674 gold 318 silver 993 10101 ToU.S.mini

FIG. ,79. Flow sheet of the concentrate treatment plant at Idaho-Maryland Mines, Grass Valley, California. CiJ CiJ

co

340


and carrying small amounts of pyrite, marcasite, chalcopyrite, and galena. Two-stage grinding, with the classifier overflow from the primary-mill circuit passing to three secondary mills, has recently been adopted in this plant to produce a finer cyanidation feed. The capacity of the plant is limited by the presence of slow-settling slime, and though the flow sheet is that of a standard C.C.D. layout, the scheme of overflowing the primary thickeners (low-lime stage) at 5 per cent solids, by-passing the agitators, and putting this overflow into the first thickener of the C.C.D. series (high lime) has made it possible to maintain tonnage without lowering extraction. Cyanide is added at the first agitator and lime at the second agitator of the series. A point of interest was the construction of a milk-of-lime plant at this property for a total cost of $10,000. The lime, which is shipped in 20 miles by plane and costs $40 per ton, is conveyed by a belt feeder from a receiving bin to a small Denver ball mill closed-circuited with a locally constructed Dorr-type classifier. The overflow passes to a Denver conditioner, from which point it is distributed to the plant. La Luz Mines, Ltd. (Type IVe) is running a plant at Siuna which is handling 800 tons of mine are and 550 tons of surface are per day, but on completion of a washing plant for the surface are, it is expected to increase the tonnage to 1800 to 2000 tons per day. In this plant the minus 48-mesh fines will by-pass the fine-crushing section and pass directly to flotation. From the fine-ore bins the ore is ground in three 8- by 8-ft. AllisChalmers rod mills closed-circuited with three 8- by 26-ft. by 8-in. Dorr classifiers. Gold traps are installed in the mill-classifier circuit, the concentrate from which is periodically cleaned on a Wilfley table and sent to the refinery. The classifier overflow is then floated in three six-cell Denver flotation machines (roughers) and two Southwestern machines (scavengers), the flotation tailings, together with the cyanide residues, passing over forty-eight 4- by 6-ft. corduroy tables. All concentrates are next classified in a 48-in. Denver spiral classifier, the sands being reground in a 36-in. by 8-ft. Hardinge mill closed-circuited with a 4- by 24ft. Dorr classifier before passing to the cyanide circuit, while the overflow after thickening is sent to the second agitator of the series. (The thickener overflow, which is usually slimy, passes to the second cyanide thickener.) The cyanide circuit comprises three 20- by 20-ft. Dorr agitators and three 38-ft. DOlT thickeners with two agitation steps between the first and second thickener. Some 30 per cent gold recovery is made in the hydraulic traps, and 60 per cent by cyaniclation-90 per cent total. The reagent consumption is given in Table 59. The Aerofloat and a part of the pine oil are added to the ball mills. The ratio of concentrations by flotation is 30: 1. The mills discharge at 75 to 80 per cent solids, and

341


classifiers overflow at 30 to 32 pel' cent solids. The circulating load is 300 pel' cent. The calculated grinding rate is 1.56 hp. pel' ton of minus200-mesh material produced in the mill. The presence of copper in the Fine ore bins

---r-

.---~ 3-a'x a' Allis Chalmers boll mills

+

4-Gold trofs in circUiT - - Conc ~

3-8'X 26'X a" Dorr clossifiers

u: 10

WIYfley toble

(Sotds) (OvefflolW 1..-----.,;-3-J_6 cell Denfer flot cells (Tarls) (tone)

(Tats)

Refinery"

2-South-westerj flot machines

(etc.)

(JIS)

.---------.......

4a-Cordufl?J::: tables

r------1~--- OJls)

(cotc.J-+l-~-----''------'

. 48"Denver sl2/ral classifier

f

f

(Sands)

(Overflow)

36" x a" H"Jding mIll 20xt'thickener closed-circuited wITh f f 4 'x 24' Dorr classifier (Underflow) (Overflow)

f f

No.1 20 'x20'Dorr agitator

-

.----+---- No. I 3a~7f2" x 12'thickener

t"

.

NO.2 8 No.3 20 x 20 DOff agtlator

Solution to precipITation

t

NO. 2 thickener (As above) -

Bullion

[

t t 2- CorduroY.. tables

Overflow

No.3 thickener (As above)

z:

(cotc.)

FIG. SO. Flow sheet of the La Luz mill at Siuna, Nicaragua.

ore has presented a problem. It is necessary to discard 500 tons of barren solution per day, and the bullion is only 550 fine, but cyanide consumption is low, and a high extraction is maintained.

S42

CYANIDA.TION AND CONCENTRATION OF ORES HONDURAS 14

The mineral districts in this country are widely scattered, and because transportation is so difficult, few deposits are sufficiently rich to justify mining development. Two mining enterprises furnish practically the entire mineral production of Honduras today. These are the N ew York and Honduras Rosario Mining Co. at San Juancito, 100 miles from the Pacific Coast, and the Compania Minera Aqua Fria, near Danli, about the same distance from the coast but near the Nicaraguan border. The Rosario mill of the former company, in continuous operation for nearly 70 years, has a present milling capacity of 550 tons per day and had produced up to 1942 more than 62 million dollars in gold and silver bullion. In the single year 1941, nearly 3},~ million ounces of silver and 23,000 oz. of gold were recovered from the ore mined. The recently constructed Mochito mill, operated by the same company, is located near Lake Yocjoa and is milling 100 tons per day of a high-grade silver ore (see Chap. XVI). TABLE

59.

REAGENT CONSUMPTION AT LA Luz

Flotation plant: Na xanthate Aerofioat 25 Pine oil.. Cyanide plant: Zinc dust Lead acetate

0.028 lb. pel' ton 0.065 " 0.034 0.020 0.005

SOUTH AMERICA COLOMBIA1 5

This state occupies first place in the production of gold in South America, and the industry is largely the basis of the country's economy. Silver and platinum, which is extracted mainly in the rich region of EI Choco, are both by-products of the gold-mining industry. In the eight principal departments, or districts, reporting gold production up to 1880, Antioquia and Cauca were first place with a total of 100 million pounds sterling. In the Choco district between 1861 and 1928, some 2300 precious-metal mines were proclaimed of which 25 per cent were vein deposits. Current gold production is obtained from lode, dredging, and hydraulic operations. The lode gold occurs, as in other parts of the world, chiefly in shoots, and continuous strikes of payable values are exceptional. However, in 14 Partly abstracted from an article by K. H. Matheson, E. and M. J., Vol. 143, No.8, August, 1942. 15 Mining Muq., December, 1946; E. and M. J., Vol. 143, No.4, April, May, and June, 1942; and also article by A. E. Villa in August, 1942; issue of same volume.


343

the auro-argentiferous type of mineralization it is quite common for the silver and base-metal values to persist throughout, although' the gold itself is localized in shoots. Blende and tetrahedrite in unusual concentrations often herald exceptional gold values. In general, however. it is the opinion of the author of the article that the size of the deposits have not justified large-scale mining operations. One of the largest gold producers is the Frontino Gold Mines, Ltd., at Segovia, where high-grade ore is being milled in a 45-stamp, sand-andslimes plant at the Silenco mine and a 30-stamp mill and cyanide plant for the Marmajito mine. The output for 1945 was 97,320 tons milled with a yield of 51,712 oz. gold and 47,061 oz. sii~ler. TABLE

60.

1941 SUMMARY

OF OUTPUT FROM VEIN AND ALLUVIAL GOLD MINES IN COLOMBIA

Type of mill or appliance

Percent Oz. gold distriper month bution

No.of No. of mills stamps

Antioquian mills ...... California mills ....... Ball mills ........ '..... Drag mills ............

460 68 12 400

Dredges ............-... Draglines ............. Hydraulic elevators ... Monitors ..............

17 2 110 400

2300 480

1610 tons 1300 tons 500 tons 50 tons

per per per per

day day day day

3460 tons per day 3,100,000 yd. per mo. 60,000 yd. per mo. 880; 000 yd. per mo. 20,000 yd. per mo.

25,000

49

26,100

51

4,060,000

51,100

--

Total. .............. ...... .... ..

The Pato Consolidated Dredging Company is the largest of the alluvial operators :in Colombia. The company operates five dredges and five hydraulicing plants, with a total capacity of 17 million cubic yards per annum, and controls practically all the' gold-bearing ground along 16 miles of the Nechi River. The Timmins Ochali Mill (Type VC)16 is a 300-ton, all-sliming cyanide plant. It is situated near Yaryumal and is built on a steep hillside to take advantage of gravity flow. A heavy rainfall ensures an adequate supply of water. . Crushing ancl Grincling. The crushing plant crushes 22 tons of minus lO-in. mine ore to 40 per cent m:inus 7'4 in. each hour. Grinding is accomplished in three 5- by 8-ft. ball mills in closed circuit with spiral classi10 Forbes K. Wilson and Balfour F. Darnell, "A Lode Gold Mine in Colombia," E. and 111. J., Vol. 143, pp. 58-61, May, 1942.

344


fiers. All cyanide and lime required in the circuit, except for the concentrate circuit, are added at the ball mills, and 68 per cent extraction is made before the pulp leaves the ball-mill classifier circuit. Gravity Concentration and Regrind. Each regrinding circuit product is split to pass over three double-deck Deister slime tables in parallel. The table concentrates are pumped to a regrind ball-mill classifier circuit, where additional lime and cyanide are added. The slimed concentrate is thickened in a 15-ft. DOlT thickener -and flows next to three parallel banks of three Pachuca agitators. Their discharge is sent to primary slime agitation. Thickening, Agitation, and Filtration. The slime product of the Deister tables is thickened in three DOlT thickeners, the overflow being delivered directly to precipitation. The thickened slimes and the discharge of the concentrate agitators flow, to three Pachuca primary slime agitators in series. The slime-agitator discharge is filtered on drum filters, the filter cake being repulped with barren solution in a paddle-type repulper. Pachuoa agitation and filtration are repeated for a second washing. Precipitation. Only the overflow from the slime thickeners is used as feed to the Merrill-Crowe simultaneous clarification, de-aeration, and precipitation plant. Three 200-ton precipitation units consume 0.23 lb zinc dust per ounce of fine gold precipitated, and 1,700 lb. raw, wet precipitate is produced monthly. Refining. Precipitate is dried by vacuum for easy handling and is treated in batches in the acid plant. For each pound (dry weight) of precipitate, 2 lb. HCl is charged into the acid tank. After agitation with steam for half an hour.. the acid liquor is decanted. The sludge is washed with hot water and then forced by ail' pressure into the Shriver press, where it is washed again. The press cake is finally dried to 6 per cent moisture in an electric oven. A typical charge to the melting furnace consists of 57 per cent precipitate, 16 per cent borax glass, 12 per cent silica, 11 per cent manganese dioxide, 4 per cent soda, and 1 lb. niter. A typical bullion bar assays 710 fine gold and 270 fine silver. ECUADOR17

Mining is said to have commenced in. this state before the Inca conquest, which occurred around 1470, and was actively continued by both the Inca owners and the Spanish conquistadores succeeding them. A gradual decline later brought mining to a virtual standstill. This was the result of a number of factors to which poor transportation, political instability, and high interest rates on loans all contributed. 17 Partly abstracted from an article by H. W. Van Putte, E. and 111. J., Vol. 143, No.8, August, 1942.


345

A moderate revival of mining was brought about some 50 years ago when the South American Development Co. brought into production the Portovelo gold mine in the southern part of the republic. Later the lVIacuchi mine, since closed down, was operated by the Cotopaxi Exploration Co., a subsidiary of the above company, and became' the second largest producer of gold in the country. A high-grade copper-gold concentrate was sintered and later smelted to a blister copper that was shipped to United States refineries. Apart from a few small washing plants, there are no other gold producers of importance in Ecuador today, although certain areas of the country are known to be highly mineralized. J Mill Lime

I'----~ •

0'( bins

50-9001b staml2.s

.

,~-

4-6 x 14 Marcy mills closedcircuited with Dorr classifiers

NaCN

l

t

2-Dewatertn~ thickeners

~

---i. .

l

--r4-Dorr aoltators

Soluble solts to waste

3- SecondarY....!.t.'fl,'h/rf'cklfJ.e~n'§er.~s:..- - - - - -

l.

t

2 Sta9..es of filiration

Pregnant solution' clarification

-(Filter cokef reputped)

t

~r_.:Fi:10Jt[t:io=n:..-__.. I

_______ J:'

. "

t

. Galena c,!once;trote

OopperccJncentrote'

t------l

'f

Tailings

f

To waste

Merrill - Crowe I2.recinilotion .

--T,

Gold bullion

~ed 10smelter

FiG. 81. Flow sheet of the Portovelo mill in Ecuador.

Portovelo Mill (Type Ills). The 350-ton mill of the South American Development Co. at Portovelo is treating a complex gold ore carrying chalcopyrite, galena, sphalerite, and pyrite. Considerable changes have been made in the original flow sheet. The old scheme of contacting the crushed ore with mercury' in revolving cylinders has been abandoned, and a modern cyanidation flow sheet (Fig. 81), followed by flotation to recover base metals, has been adopted. About 410 tons of ore is hoisted per day, but some GO tons of waste is sorted out at 1p,;! in. after the crushed rock has passed through washing trommels. Mter sorting, the mill feed assays per ton 0.43 oz. Au, 2.0 oz. Ag, O.GO per cent Cu, and 0.4 per cent Pb.

346 , The


U~-in.

feed is crushed in batteries of 900-lb. stamps fitted with screens, and the discharge ground in Marcy mills in a water circuit with the addition of lime to the mills and using a composite load consisting of 15 per cent l>~-in. balls and 85 per cent 4-in. rhyolite pebbles. The 75 per cent minus-200-mesh pulp is then thickened, and the clear overflow, which carries soluble sulphates, is discarded. Agitation is carried out for a period of 28 hr. on the thicker underflow at 35 per cent solids, cyanide being added to the first agitator. Following secondary thickening and two stages of filtration on drum filters, with a water wash on the second stage, the cyanide residues are repulped and floated for the recovery of the base metals. A xanthatecresylic acid circuit is used for this purpose, and about 35 tons per month of a galena concentrate carrying 100 oz. silver per ton is shipped to the Selby smelter in California, and 100 tons of copper concentrate pel' month shipped to Carteret, N.J. On account of high transportation cost, the grade of these concentrates is apparently more important than the metal recovery. About 1200 to 1400 tons per day of pregnant solutions coming off the secondary thickeners and the filters is clarified and then precipitated by the Merrill-Crewe process. A total recovery of 93.5 per cent of the gold and 55 per cent of the silver in the mill heads is made in this plant, some 20 per cent of the silver recovery being made by flotation. The ore is ground to 75 per cent minus 200 mesh, and 0.9 lb. N aCN and 0.1 to 0.2 lb. CaO per ton solution strength is maintained. The cyanide consumption is 2.2 lb., and lime consumption is 6.0 lb. per ton of ore fed to the mill. The lime is burned locally, using producer gas. Calera Exploration Co. The Portovelo mill is also handling 100 tons per day of Calera ore, which, although of the same general mineral content, requires a different method of treatment, and a separate circuit is employed for the purpose. The classifier overflow, instead of passing directly to thickeners, is hydroseparated to produce a slime overflow that carries a high proportion of soluble copper and can either be by-passed to the copper flotation circuit or be separately agitated in a dilute cyanide circuit to cut down the copper dissolution. The gold precipitate will run as high as 30 per cent copper. The 'hydroseparator sands are first agitated for 8 hr. and then passed to a primary thickener, where pregnant solution is withdrawn, The thickened pulp is then further agitated for 28 hr. in a series of four agitators and thickened again before stage filtration. The interesting feature of this circuit is that one single-tray closedtype thickener is used for both the first- and second-stage thickening. The residues are floated for production of a lead and a copper concen~16-in.


trate, as in the case of the Portovelo ore, and grades up to 65 pel' cent Pb and 22 pel' cent Cu are made. The Calera heads after sorting assay 0.54 oz. Au and 2.0 oz. Ag pel' ton, and a 6: 1 precipitation ratio is maintained. BRAZIL

Gold" is found in most of the Brazilian states, Minas Geraes being the largest producer. In the central part of the state we find the famous Morro Velho mine of the St. John Del Rey Mining Company, one of the deepest mines in the world. Smaller deposits of low-grade ore are being worked neal' Ouro Prete, Caete, and Santa Barbara, and in limited areas in Parana and Rio Grande do SuI. Prospects for new developments are encouraging wherever the Minas series schists are found. In Minas Geraes, Goyas, Matto Grosso, and Maranhao conditions are favorable and many good prospects are already being studied. The Passagem gold mine is about 4 miles east of Ouro Prete. The company also operates the Santana mine near Passagem. Production is approximately 300 tons. It ranks second in Brazil in importance and, like Morro Velho, is very old. Gold is irregularly distributed throughout; the quartz associated with tourmaline, pyrrhotite, and arsenopyrite. Milling practice follows closely that of the other gold mines in the area. It includes two-stage crushing, grinding, concentration on jigs and tables, and cyaniding resultant concentrates. There is also a furnace for the recovery of arsenic. The Juca Vieira gold mine is situated neal' the town of Caete about 53 kilometers from Bello Horizonte. About 150 tons per day' of a complex are, resembling that at Morro Velho, is mined. The mill includes jigs in the grinding circuit and flotation equipment which makes a concentrate that is cyanided after grinding to 300 mesh. Tailings from the flotation machines are classified and passed over a pulsator jig and concentrating table, the gravity concentrates from tliese machines being returned to the head of the mill. A total recovery of 97.4 pel' cent is made. Operating costs are $3.75 per ton. Milling Practice at St. John Del Rey Mining Company Ltd. Ore' is received from two main sources, the Morro Velho mine and the Raposos-Espirito Santo mine, and is treated in a central mill. The important minerals in Morro Velho ore are 20 pel' cent silica, 32 per cent iron, 15 pel' cent sulphur, 3}/~ pel' cent arsenic, 18 per cent lime, and 7H per cent magnesia. It consists of a finely crystalline admixture 18 Abstracted from articles by John B. Huttl and E. A. Teixeira, E. and M. J., Vol. 143, Nos. 3 and 8.

348


of quartz, calcite, dolomite, siderite, ankerite, pyrrhotite, pyrite, and arsenopyrite with some chalcopyrite and traces of other metals. Pyrite and arsenopyrite are the only minerals present in any definite crystalline form as visible to the naked eye, whereas the other minerals occur as a fine-grained mass, chalcopyrite having a tendency to segregate. Pyrrhotite is the predominant sulphide, readily recognizable because it rapidly tarnishes to all shades of steel blue, pinkish red, and yellowish brown. Chalmersite, a copper pyrrhotite, was first recognized here. The gold content of the ore averages 13 grams per metric ton, and the silver content about 3 grams. The metals occur in a finely divided condition as a native alloy of 77 per cent gold and 23 per cent silver. Visible gold is rare. Microscopic examination of polished surfaces of the ore under reflected light reveals that gold is most generally associated with arsenopyrite and in contact with pyrrhotite, similar to the relation of gold to pyrite and pyrrhotite mentioned by Paige in Bul, 765, U.S.G.S., 1924. Raposos-Espirito Santo ore, while carrying the same minerals as those occurring in Morro Velho ore, is much more siliceous, and its sulphur and arsenic contents are considerably lower. It averages approximately 60 per cent silica, 6 per cent sulphur, and 1 per cent arsenic, and the gold content is about 8 grams per ton. Ore treatment comprises crushing, fine grinding, and concentration followed by cyanidation, and until this last step is reached, treatment of the two ores, although carried out in the same mill, is kept separate. Details of the individual operations follow. Fine Grinding and Concentration, lI1~orTO T1elho Ore. The ore is reduced to approximately 5 in. in a primary jaw breaker. Grizzly undersize and crusher product pass through a trommel delivering 2-in. ore to secondary jaw crushers, which make a 2-in. product. One hundred fifty Californian stamps crush the ore to 30 mesh with a stamp duty of 6 tons per day. Stamp-mill product is roughly classified in spigot launders, the underflow from which is concentrated on James sand tables. Table tailings are dewatered in automatic cone classifiers and fine-ground in tube mills working in closed circuit with James sand tables and the cone classifiers to approximately 70 per cent minus 325 mesh. Cone classifier overflow, together with spigot-launder overflow, is thickened and concentrated on Holman-James sand tables. Thickener overflow goes to cyanide, while the Holman-James table tailings are given a final scavenging in Andrews classifiers, the small amount of coarse material from which is returned to the tube-mill circuit while the overflow goes to cyanide.


349

Raposos-Espirito Santo Ore. Primary crushing is carried out underground in a jaw crusher. Surface plant consisting of two Symons cone crushers in series, working in closed circuit with Hummer screens, reduces all the ore to minus % in. An aerial ropeway transports the crushed ore . to the central mill. A Hardinge conical ball mill, in closed circuit with a Dorr classifier, is used for primary grinding. Classifier overflow feeds spigot launders whose underflow is concentrated on James sand tables. Both spigot-launder overflows and table tailings are classified in two Dorr bowl classifiers working in parallel and delivering a sand productwhich is fine-ground in a Hardinge mill running in closed circuit with tables and the bowl classifiers. Classifier overflow, averaging about 70 pel' cent minus 325 mesh, is cyanided. All tables in the mill, whether treating Morro Velho or Raposos-Espirito Santo ore, make three products: concentrate, middling, and tailing. Concentrate Treatment. Concentrate from all tables, consisting of free gold and sulphides, practically wholly arsenopyrite, is combined and recleaned on one Holman-James sand table. This table produces a concentrate which is smelted and accounts for approximately 50 per cent of the total gold recovery. Table tailing joins the middling circuit. _M~iddling Treatment. Table middlings consist of arsenopyrite and pyrrhotite with lesser amounts of pyrite and gangue minerals. They, together with the tailing from the concentrate table, are dewatered in a Dorr classifier and ground to approximately 85 per cent minus 325 mesh in a cylindrical ball mill working in closed circuit with two Holman-James sand tables and a Dorr hydroseparator. Overflow from the hydroseparator is cyanided. Cyanidation. Morro Velho circuit tailings are dewatered in a 120-ft. Dorr traction thickener, and Raposos-Espirito Santo circuit tailings in a 50-ft. thickener. Thickener underflows are combined and cyanided in four 12-ft.-deep by 25-ft.-cliameter agitators in series. Cyanide, lead nitrate, and lime are added at beginning of cyanidation, and aeration is applied vigorously. Treatment time is 5 to 7 hr., and cyanided pulp is filtered on Oliver filters. Filter cake is pumped to storage in dams, while filtrate is clarified and precipitated in a standard Merrill-Crewe Plant. The finely ground middlings are thickened in a 30-ft. Dorr thickener and agitated for 10 to 12 hr. with cyanide, lead nitrate, and lime in three 5-ft.-deep by 17-ft. 5-in.-diameter agitators. Extremely thorough aeration has been found to be necessary for successful cyanidation of this material. Flotation. Following cyanidation, middlings are washed in a four-com-

350


partment Dorr tray thickener, overflow from which is clarified mid precipitated in a standard Merrill-Crewe plant, while the underflow is filtered on an Oliver filter. Filter cake is floated with pine oil, xanthate, and copper sulphate to concentrate the gold and arsenic values in as small a bulk as possible. The concentrate, whi'ch averages about 25 pel' cent arsenic and 10 grams of gold pel' ton, is roasted, while flotation tailing, consisting mainly of pyrrhotite, goes to waste. Roasting is carried out in a standard Edwards roaster, and volatilized arsenic is precipitated in a Lodge-Cottrell plant. Calcines are returned to the mill, ground in a tube mill loaded with quartz pebbles, washed in a 3D-ft. thickener, and cyanided for 3G hr. with cyanide and lime. Crude arsenic from the Cottrell plant is refined in reverberatory furnaces for marketing. PERU19

Some 30 per cent of the gold produced in Peru is a by-product of the copper, lead, and zinc industries. The remaining 70 per cent comes from properties which produce only gold and which belong to companies owned by Peruvian and American capital; hence the extraction of gold in Peru is the only important industry which can be consi'dered nationally owned. It employs about 9000 men, or about 33 per cent of the total mining laborers. Placer mining represents about 9 per cent of the total production. The greater proportion of gold mining in Peru has been carried out on the coastal plains, but these mines are really small and offer few possibilities for large-scale operations. The real gold future of Peru is in the Montana, or jungle, region east of the Andean ranges, where most of the rivers carry gold in commercial quantities and where a considerable number of terraces exist that were worked extensively and profitably by the colonial Spaniards and still contain large volumes of gold-bearing gravels. Peru is the world's fourth largest producer of silver. About GO per cent of the total production is a by-product of lead- and copper-ore reduction, so that the production of silver is in direct proportion to the extraction of lead, copper, and zinc ores. Silver is widely spread but in most cases represents the basis of small enterprises. VENEZUELA2 0

Geographically and geologically Venezuela is roughly divided into two sections, one north and the other south of the Orinoco River. The eastern half of the southern section forms the state of Bolivia. This region is also known as the Venezulan Guayana (01' Guiana). 1D

20

Article by Carlos DelSolar, E. and 111J., Vol. 143, No.8, August, 1942. Article by N. B. Knox, E. and 111J., Vol. 143, No.3, March, 1942.


351

The most important gold mineralization is centered at El Callao, a district which in the 1860's and 1870's was one of the world's leading producing regions. The El Callao mine itself alone produced 175 million gold francs from 3-oz. rock treated in a 60-stamp mill. This camp is situated about 200 kilometers airline southeast of Ciudad Bolivar, state capital on the Orinoco. About 95,000 oz. of gold is produced annually from this district out of a total for Venezuela of 114,000 oz. (1938), but aside from the metal obtained from alluvial washing with the batea, the only production in the district is from two hard-rock mining companies; the Moeupia, a French company, and the New Goldfields of Venezuela, Ltd., which is British. These companies have their properties a mile or two south of the town of El Callao in a series of rocks of strikingly different aspect from the gneisses of the basement complex. A great number of quartz veins outcrop, and all of them contain some gold. A few have rich ore shoots and are mined and held as reserved. Aside from quartz and gold, the only other minerals present are calcite, ankerite, tourmaline, pyrite, and rare grains of chalcopyrite. One mill handles the ore from all the New Goldfields mines. It has a capacity of about 600 tons pel' day. Treatment is by a combination of cyanidation and flotation. Extraction is 94 per cent, working on l l-dwt. heads. About 8000 oz. gold are recovered from 17,000 to 18,000 tons of ore pel' month. The Mocupia Mining Company is working the Colombia lode, which is within an enclave in the properties of the New Goldfields Company. The vein is similar in structure and mineralization to those of its neighbor and is cut bya norite dike. The mill makes about 70 per cent recovery. ClIILE 21

Gold has been mined in this country without interruption since the days of the conquest and even before that time by the natives, but only after 1932, following the crash in 1929, did gold mining really become profitable. Practically all the gold veins so far found are within the Coast Range or in the longitudinal valley which runs north-south practically the whole length of the country. Production of gold which had remained at around 32,000 oz. pel' year, mostly derived from copper ores and concentrates, has risen in 10 years to over 320,000 oz. per year. Though many thousands of gold mines have been discovered and worked in the last 10 years, the number of those which have developed from small prospects into really good-sized mines has been remarkably few. Among these are the rich veins of the Altamira district near Taltal, the rich gold21

Article by Fernando Benitez, E. and M.J., Vol. 143, No.8, August, 1942.

352


copper veins of La Isla at Inca de Oro, the remarkable Capote vein which wae worked by the Indians, the new mine at Tres Amantes, and the lowgrade deposits of Andacollo where occasionally very high-grade stringers are found containing gold in the form of ribbon or wire. (In this field two flotation mills handling 450 tons per day have been built.) At Bellaviste there is a deposit carrying gold copper and zinc, but the most important gold mine opened in Chile in recent years is the Punitaqui near the town of Ovalle where the are carries 0.25 oz. of gold, some copper, and appreciable values in mercury. The ore is concentrated by flotation in a mill which handles 400 tons per day. The mercury are is treated in one of the two sections of the mill (each of 200 tons capacity) quite apart from the gold-copper are, and a flotation concentrate made which runs 20 per cent mercury and 20 grams gold to the ton. This is distilled in the usual cast-iron retorts, which are oil-fired. Production for the second semester of 1941 amounted to 10,700 oz. gold, 278 tons copper, and 1280 flasks of mercury. BRITISH GUIANA

Southeast of Venezuela lie the Guianas (British, Dutch, and French). The British Guiana Consolidated Goldfields, Ltd., operate an electric bucket dredge on the Upper Mahdia River and during the year ending July 31, 1946, recovered 8,042 oz. gold from 1,172,880 cu. yd. of material at a cost of 8A5d per cubic yard. Large gold-bearing areas are known to exist in this country, but the recovery of the gold presents considerable difficulties owing to the high clay content of the oxidized ore and "spotty" nature of the values present.

SECTION 3.

AFRICA

SOUTH AFRICA-WITWATERSRAND

Ore mined in the vicinity of Johannesburg, Transvaal, is a conglomerate composed of various pebbles, principally of barren white quartz, which constitutes 70 per cent of the whole. The pebbles are within a matrix of cementing material consisting of fine-grained quartz, sericite, chlorite, and chloritoid. The matrix carries the gold, a little silver, and the mineral sulphides. Argillaceous components of the mill pulp may be as high as 32 per cent. Of the sulphides, pyrite and pyrrhotite predominate. Some osmiridium is also present-v-I oz. in 3000 to 10,000 tons are. Most of the oxidized ore has been extracted, so milling is confined mainly to the sulphide are. The average value of Rand are (47 plants) is 3.95 dwt. per ton. After waste averaging 0.188 dwt. per ton has been sorted out, the are treated is


353

about 4.25 dwt. per ton (the highest head value reported is 17.2 dwt; per ton). . GENERAL PRACTICE

In a study of 47 plants, which hancUe close to 9:'5 per cent of the total Rand tonnage and 90 per cent of which are controlled by five principal mining groups, the following general observations can be made regarding South Mrica practice: Crushing. Jaw crushers are almost exclusively used for primary crushing down to 4 to 6 in. Both high-speed gyratory and standard co~e crushers are used for secondary reduction to 17~ or 2 in., although the latter seems to be gaining in popularity. In the mills crushing in three stages the short-head cone is exclusively used for the final reduction to 7~ or % in., operating either in closed circuit or with crusher product joining screen undersize to mill bins. Closedcircuit operations are becoming more general. Screening. For the coarser sizing in the primary crushing circuit fixed grizzlies are extensively used but the tendency is toward heavy-duty vibrating screens for this duty, because of their greater efficiency, lower head loss, and reduced labor requirement. Vibrating screens are almost exclusively used in the secondary and tertiary crushing circuits. Washing. In most plants plus 1 to 1~ in. is washed, and washing all ore plus 7~ to % in. is becoming common practice, while in some plants the whole ore is washed. As washing is required for are to be sorted, extension of washing to all are plus ~ in. can be made with little increase in capital or operating cost. 22 There are three interrelated factors which have led to washing of are becoming a more important feature in are dressing on the Rand. 1. Conditions of wet mining and long, steep ore passes make for delivery to the metallurgical plant of large quantities of wet, sticky fines, difficult to handle on conveyor belts and in subsequent sizing operations. 2. Fine crushing to a size most suitable for feed to pebble and ball mills has introduced crushers which cannot deal with a feed carrying much wet fines. 3. In turn, the use of fine crushers has called for the introduction of vibrating screens to separate down to, say, 7~-in. mesh, and with such equipment "primary slime" reduces efficiency to an absurd level unless washing is introduced. 22 A. Clemes, "Modern Metallurgical Practice C.M. and ill.S.S.A., Vol. 48, No.2, August, 1947.

III

the Witwatersrand," Jour.

354


61.

WITWATERSRAND GOLD PRODUCERS

1646

1947

FiOperating company

Blyvooruitaicht G. M. Co., Ltd.. .......... Bmkpun Mines. Ltd .... City Deep, Ltd......... Cons. Main Reef Mines and Estate, Ltd...... Crown Mines. Ltd ..... . Daggufontein Mines, Ltd ................... Roodepoort Durban Deep Mines, Ltd ..... East Champ d'Or G. M, Co., Ltd.......... East Dnggafntein Mines, Ltd........... Eust Geduld Minos, Ltd................... East Rand Proprietary Mines, Ltd........... Proprietary Geduld Mines, Ltd........... Government G. M. Areas Cons., Ltd..... Grootvlei Proprietary Mines, Ltd........... Luipuards Vlei G. M, Co., Ltd .............. Marievule Cons. Mines, Ltd................... Modderfontein East, Ltd................... New Kleinfontein ...... New Modderfontein G. M, Co" Ltd .......... New States Areas. Ltd. Nigel G. M, Co., Ltd... Nourse Mines, Ltd ..... Randfontein Estates G. M. Co .... ......... Rand Leases G, M. Co., Ltd.. ........... Rietfontein Cons. Mines. Ltd........... Robinson Deep, Ltd... Rose Deep, Ltd. ...... Simmer and Juck Mines, Ltd... ........ So. African Lund und Exploration Co.. . . . . . Spring Mines, Ltd ...... Sub-Nigel G. M. Co., Ltd................... Van Dyk Cons. Mines, Ltd, ..... , ............

Dunce group

Dry tons milled

Yield Yield Working Fine Fine Working per Dry tons per ounces cost ounces cost ton, milled ton, produced pel' ton produced per ton dwt, dwt. -------

---

C A C

285,600 15.930 227,486 1,257,OUU 4.UOI 251,486 996,000 5.113 254,652

60s 4d 30 U 36 4

459,00U 16.679 1,206,UUU 3.945 965,00U 5.126

382,777 237,913 247,,1,76

46s Od 32 3 38 5

C C

2,464,000 3,16U,000

2.723 236,520 '1. 273 675,159

21 2 26 5

2,191,000 3,426,000

3.058 3.520

335,024 602,661

24 3 2610

A

1,874,00U

5.479

513,428

20 3

1,746,000

5.353

467,341

2011

C

2,036,000

3.716

378,281

25 9

1,688,000

3.767

374,476

2510

J

326,000

3.845

63,250

27 2

322,000

3.697

61,374

28 6

A

1,085,000

5.057

274,317

25 7

645,500

4.726

223,582

28 3

U

1,862,000

5.633

524,445

18 0

1,722,000

5.482

472,023

19 0

C

2,443,000

3.661

487,520

27 5

2,525,000

4.408

556,464

27 2

U

1,265,000

4.234

267,787

23 11

1,167,00U

3.657

236,847

2310

J

2,564,000

3.666

475,877

27 2

2,544,000

3.524

4'18,288

27 4

U

1,622,000

5.150

464,602

21 11

1,785,OUO

5.015

447,572

23 8

G

969,500

4,134

206,575

2810

968,5UO

4.055

166,385

26 5

U

649,000

6.006

164,963

37 2

617,000

5.626

182,825

37 7

C

1,517,000 1,166,500

2.662 3.055

226,941 182,745

1711 23 1

1,386,000 1,200,00U

2.820 3,014

195,826 180,820

18 9 23 2

C J

2.661 4,157 5.114 3.7'-4

141,916

C

1,029,000 1,321,00U 489,000 836,000

125,043 158,609

22 22 37 30

6 6 7 8

872,000 1,249,000 450,400 846,000

2.642 3.763 4.325 3.663

120,206 234,996 67,'10·[ 154,654

22 25 37 30

J

4,113,000

2.541

522,455

16 6

3,781,000

2.557

483,381

2011

T

2,172,000

3.787

411,239

25 8

2,141,000

3.366

360,302

26 1

G G C

315,000 1,141,OUO 853,000

64,691 4.107 3.614 223,266 3.U16 128,635

23 5 31 6 24 8

262,000 1,252.000 838.000

4.245 3.755 2.953

61,677 235,086 123,711

26 5 30 7 25 6

G

1,564,000

3.638

289,916

27 6

1,427,00U

3.639

256,669

29 10

A A

1,052,500 1,437,000

4.007 3.646

210,882 261,978

25 8 27 4

1,028,500 1,351,500

3.611 3.365

201,116 226,404

26 3 28 5

760,000 10.143

400,667

37 8

781,500

6.665

377,6'16

38 3

3,618

232,717

28 6

1,145,30U

3.713

212,636

26 4

G U

1,188,000

247,592

2 7 1 3


61.

WITWATERSRAND GOLD PRODUCERS

355 (Conlinned)

1946

1947

Fi-

Opcru.ting company

Yield' nunce Working Fine pel' group Dry tons cost ounces ton, milled produced per ton

Dry tons milled

dwt.

---

Vnn Ryn G. M. Estate, Ltd ... Venterspost G. M. Co., Ltd .... ... . . . . . . . . . . Vlakfontein G. M. Co., ...... ...... Ltd.... Vcgelst.ru iahul t G. M. Areas Co ... ... ...... Western Reefs Exp. and Dev. Co., Ltd... West Rand Cons. G. ...... M., Ltd ... West Springs, Ltd .. Witwatersrand G. M. Co., Ltd.. ........... Witwutersrand Nigel, .... ..... Ltd ... Miscellaneous ....... .. Totals and A vemges .

-~

--- ---

M

689,000

2.304

79,367

19s 8d

G

1,277,500

4.160

265,725

Yield Fine Working per .ounces cost ton, produced per ton dwt. --- ---

534,000

2.369

64,043

19s 9d

28 1

1,157,000

4.175

241,528

30 4

8.175

111,383

56 0

G

27'1,000 8.383

IB,849

53 9

., 272,000

G

886,000

4.9B

217,692

30 9

789,000

4.972

190,162

3,j10

A

914,000

4.796

219,178

28 4

912,500

4.918

224,370

29 2

M A

2,580,000 695,500

3.625 4.048

467,581 BO,773

21 1 30 8

2,369,000 701,000

3.412 3.575

404,122 125,292

2210 30 1

949,000

2.703

128,256

22 3

870,000

2.682

116,677

22 7

29,2,j8 682,339

52 11

J

48 ,j

T

106,300 5.767 30,649 105,800 5.529 794,865 .. . .... 4,000,000' . 3,800,000' . ..... -- --- -- --- --- --- -58, 7~7,500

4.024 1l,m7,914

'" Estimated. C-Central Mining and Investment Corp. A-Anglo-American' Corp. of S. A., Ltd. T-Anglo-Tl'unsvuul Cons. In,~estment Co., Ltd. G-New Cons. Gold Fields, Ltd.

57,512,300

---

3.982 11,197,638

J-Johunnesbul'g Cons. Investment Co., Ltd. l\'I-Generul Mining and Finance Corp., Ltd. U-Union Corp., Ltd.

'Washing is done on vibrating screens, stationary grids, individual washing belts, or the lower end of the sorting belts. The washing fines (usually minus 7~ in.) are classified, with rake product conveyed to mill bins and overflow (minus 100 mesh) thickened and pumped to the grinding circuit or to the cyanide plant. Waste Sorting. Sorting of waste is generally practiced, it being included in nearly 80 per cent of the plants. In those plants doing sorting, the average amount discarded as waste is about 9 per cent of the crude are. The average value of waste is under 0.2 dwt. per ton. More generally sorting is done on only one size, usually plus 3-in., but in some plants fine sorting down to 17~-in. size is practiced. Sorting is done by natives from inclined belts, usually 42 in. wide, running at about 70 ft. per min., and on an average the natives can sort out about 1:J1 tons per hr. per man on the coarse sizes down "to less than 7'2 ton per hr. on the fine sizes. Stamping. While about 40 per cent of all Rand are is handled through Stamps, no new stamps have been installed in the last 30 years. In the older plants still using stamps, they are used for final crushing of 1- to 2-in. material to tube-mill feed.

356


Grinding. Two-stage grinding is general practice. For primary grinding-sometimes open-circuit but generally closed-circuit ball mills-composite mills (using a mixture of ore pebbles and balls) and straight pebble mills are used. For secondary grinding, composite and pebble mills are used. Most mills are of the low-discharge grate type, and pumps, rather than scoop feeders, are used for returning the classifier sands to the mills. In over 80 per cent of the plants the grinding is done in an alkaline-water circuit. The average fineness of grind in all plants is 66.7 per cent minus 200 mesh; in all slime plants 73.2 per cent minus 200 mesh. Concentration. Very little concentration other than by corduroy tables is practiced on the Rand. Corduroy tables are used in 75 per cent of the plants, and over 30 per cent of the gold recovery on the Rand is recovered on corduroy. While the gold is largely associated with the sulphides, flotation of the whole ore does not produce a discardable tailing. Two plants which do sand leaching float the sand product before cyanide leaching. Float concentrates are reground to 325 mesh and cyanided in the plant slimes circuit. Four plants use J ohnson concentrators on the whole ore. Concentrates are reground to 200 mesh and cyanided in the plant slime treatment. Sand Leaching. While many of the older plants still practice sand leaching, all plants constructed in the past 20 years are all-slime treatment. Of those plants using the dual process about 26 per cent of the tonnage is sand-treated by percolation. Only about 10 per cent of Rand ore is treated by sand leaching today. Slimes Dewatering. Intermittent thickening in plain tank collectors is still extensively used, but the tendency is definitely to adopt continuous operation in DOlT thickeners, and it is doubtful if any new plant today would use intermittent dewatering. Agitation, While less than one-third of the Rand tonnage as a whole is treated in continuous agitation, most of the newer plants employ this method. Pachuca tanks are more generally used than any other type of agitator, though a number of the newer plants use DOlT agitators. They are built in sizes up to 33 ft. in diameter by 48 ft. deep. The average period of agitation (with some exceptions) is rather short-12 to 15 hr.-· for complete dissolution. Filtering. While Butters filters are still used for nearly 25 pel' cent of the total slimes filtered, any addition to filter capacity has been made with continuous filters. and intermittent (batch) filtering would not be considered for a new installation. Clarification and Precipitation. Sand clarification tanks and Merrill-Crewe precipitation are exclusively used. From 0.25 to 0.30 ton of


357

barren solution is discarded pel' dry ton filtered. The average over-all recovery on the Rand is 95.97 pel' cent of the gold in the ore treated. Costs and Power. The average cost of treatment in 40 plants is 3s 4d (G7 cents)* per ton treated, and the average power consumption 23.8 kw-hr. per ton treated. The average nominal mill capacity is about 4000 tons per day, although one plant treats over 13,000 tons pel' day. SAND-SLIME PLANTS USING STAMPS

The mills of Randfontein Estates and West Rand Consolidated are typical of the older practice in South Africa, the former being a straightforward cyanide plant, while the latter makes a pyrite concentrate that is handled in a separate pyrite treatment plant and cyanides the rest of the ore. Randfontein Estates Gold Mining Co., Ltd. (Type IIss). The reduction plant of the Ranclfontein Estates Gold Mining Co. is the largest mill on the Witwatersrand, The nominal capacity is 13,000 tons per day. The following information and attached flow sheets were furnished by the operating staff through Fred Wartenweiler, consulting metallurgist of the Johannesburg Consolidated Investment Co. The ore is first screened on seven 8- by 3-ft. Ty-rock double-deck screens, with a 3-in. round-hole opening on tho top deck and 1~-in. square mesh screen on the bottom deck. The undersize of these screens goes directly to the mill bins, and the oversize to washing and sorting. The plus 3-in. oversize of the top deck, which ranges in size up to 14 in., is washed and sorted on five 3G-in.-wide by 118-ft.-Iong belts, from which waste and primary tube-mill pebbles are sorted. The oversize of the lower deck passes to two similar belts from which waste and secondary tube-mill pebbles are sorted after washing. The washing is done by sprays on the lower end of the sorting belts, using 525 gal. per min., and drainage from the belts or washing fines are dewatered in two simplex DOlT classifiers with rake product going to the mill bins and overflow 84 per cent minus 200 thickened in two intermittent settling tanks, from which the thickened pulp is pumped to the secondary grinding circuit classifiers. The sorting belts run at 25 ft. per min., and on an average 40 native boys pick about GO tons per hr. of waste and tube-mill pebbles. In 194G the waste amounted to 2.05 pel' cent of the crude ore and had an average value of 0.152 dwt. pel' ton. The cost was 13.0Gd (21.8 cents) per ton of waste. The sorted ore is crushed to approximately 1 in. in seven 40- by lG-in. jaw breakers set at 2;J.i in. and two 4-ft. Symons cone crushers. The

* S. A. Sterling

=

$4 basis.

FIG. 82. Flow sheet of the Rnndf'ontein Estates G. M. Co. mill, Witwatersrand.


359

crusher product, together with the primary screen undersize an~l the rake product of the washing fines classifiers, goes to 12,000-ton storage bins ahead of the stamps. The stamp battery consists of six hundred 1750-lb. stamps dropping 8Yi in. at 95 drops pel' minute. Battery screens are 7~ and % in. Stampmill feed is 30.6 pel' cent plus 1 in. and 36.6 pel' cent minus 7~ in. Discharge is 61.4 pel' cent plus 48, 11.1 pel' cent plus 100, 7.2 pel' cent plus 200, and 20.3 per cent minus 200. Grinding is done in two stages of tube mills operating in closed circuit with McLean spiral classifiers. The tube mills are grate-discharge type with ::t' 17~-in. round-hole opening on the primary mills and a %-in. round hole on the secondaries. The grinding charge is made up of pebbles and 3-in. forged balls. Primary mills carry 12 -tons of balls and 17 tons of pebbles, and secondary mills 6 tons of balls and 12 tons of pebbles. Pebble loads are maintained by the addition of 60 tons primary and 22 tons secondary pebbles pel' mill per day. Ball consumption is 0.82 lb. primary and 0.625 lb. secondary. There are eight 8- by 16-ft. primary mills operating at 21 r.p.m. driven by 400-hp. motors and eighteen 57~-in. by 22-ft. secondary mills running at 30 r.p.m., driven by 175-hp. motors. Power input is 340 hp. primary and 181 hp, secondary. Mill-discharge densities are 82.5 and 70.5 per cent solids primary and secondary, respectively. The primary mills are in closed circuit with eight 9-ft.-diameter McLean classifiers operating at 13 r.p.m, and overflowing at 76.5 pel' cent solids. The primary-mill discharge flows over thirty-two 4-ft. 4-in. by 6-ft. 3-in. corduroy tables set at a slope of 18 pel' cent. The table tails are pumped back to the primary classifiers with two 14-in. pumps driven by 360-hp. motors. The secondary mills are in closed circuit with thirty-six 7- and 8-ft.diameter McLean classifiers. operating at 9 r.p.m. and overflowing at 18 per cent solids to eighteen 7-ft. 6-in.-diameter cones. The secondarymill discharge and the cone spigot are treated on seventy-two 5- by 6-ft. corduroy tables set at a slope of 15 pel' cent. Table tails join the primaryclassifier overflow and are pumped to the secondary classifier with four 12-in. pumps driven by 175-hp. motors. The corduroy tables recover 47.7 per cent of the total gold. Corduroy concentrate amounting to about 20 tons pel' day is cleaned on Wilfley tables, and the cleaned concentrate is amalgamated and retorted. Table tails are returned to the secondary grinding circuit. The ground product (cone overflow) is pumped to six DOlT bowl classifiers 23ft. in diameter with 8 ft. wide reciprocating rakes where a sandslime separation is made.

360


Screen analysis of the grinding circuit and bowl classifiers IS given III Table 62. While the above screen gradings calculate about 30 pel' cent sand, the average of slime treated in 1946 was 79.74 pel' cent of total treated. The bowl-classifier overflow at 17.5 pel' cent solids is dewatered to 60 pel' cent solids in twenty-two 70- by 16-ft. side depth intermittent settling tanks with 12-deg. sloping bottoms. The thickened pulp is transferred to the agitators, during which cyanide is added and the pulp is diluted to 1.49 specific gravity with Butters filter weak effluent. The slimes are batch agitated for 12 to 14 hr. in fourteen Pachuca tanks 45 ft. deep with 20 ft. deep cones, twelve of which are 22 ft. in diameter and two 33 ft. in diameter. The amount of air used pel' agitator is 400 cu. ft. pel' min. at 45- to 50-lb. pressure. Cyanide added during transfer brings the initial strength up to 0.011 pel' cent KCN, and solution strength after agitation is 0.008 per cent. TABLE

62.

SIZING ANALYSIS OF GRINDING MILL PRODUCTS AT RANDFONTEIN ESTATES

Mesh

Stamp-mill discharge .. " .- ............... Primary classifier overflow ................ Secondary classifier overflow .............. Cone overflow ........ " .................. Bowl overflow ............................ Bowl sands ...............................

+48

+100

+200

-200

61.4 17.2 0.2 ....

11.1 25.5 23.4 15.0 3.7 51.8

7.2 17.9 26.0 23.8 20.1 39.7

20.3 39.4 48.6 61.2 76.2 8.5

....

....

The agitated pulp is transferred to one of two 16-ft.-deep propeller-type mechanical agitators with a 12-ft. cone bottom for storage ahead of the filters. Filtering is on Oliver and Butters filters with about 43 per cent of the tonnage handled continuously on the Olivers, There are eight 14-ft.-diameter by 16-ft.-face Oliver filters operating at 0.4 r.p.m. with one-third submergence. Duty is 1560 lb. solids pel' square foot of filter area pel' 24 hr. Barren solution at 0.34 ton pel' dry ton filtered is used for washing, and the cake is discharged at 20.6 per cent moisture. All Oliver filtrate at 0.995 dwt. per ton goes to precipitation. The Butters plant consists of 960 leaves 10 by 5 ft. The duty is 123 lb. solids pel' square foot per 24 hr. with an operating cycle of 2 hr. 10 min. (26 min. caking, 60 min. washing, and 54 min. filling and discharging). One ton of barren wash is used per dry ton filtered, and cake is discharged at 23.5 pel' cent moisture. First filtrate from the Butters at 0.995 dwt. per ton goes to precipitation, and the weak effluent is returned for dilution in the agitators.


361

The cake from both filters is diluted with barren solution and water to about 56 per cent solids and pumped to the slimes dam. . The treatment of the sands is described in Chap. VII. Pregnant solution amounting to 1.24 tons per dry ton milled'is precipitated by Merrill-Crowe vacuum-leaf precipitation system. Pregnant solu-. tion from the filters is clarified in sand clarifiers, while that from sand 63. ASSAY OF PLANT PRODUCTS AT RANDFONTEIN ESTATES Crude ore to receiving bins 2.514 dwt. gold per ton Waste sorted " 0.152 2.666, Ore to mill bins Sand to leaching. . . . . . . . . . . . . . . . . . . . . .. 1.865 1.235 Slimes to agitation Sand tails: . Washed " 0.201 Unwashed 0.214 Agitator tails: Washed 0.082 Unwashed 0.097 Filter residue: Washed , 0081 Unwashed 0.096 Pregnant solution 0.992 Barren solution 0.017 Dam solution 0.003 Oliver cake moisture 0.040 Butters cake moisture 0.050

TABLE

64. REAGENT CONSUMPTION Cyanide (KCN) Lime (CaO) Zinc Lead nitrate Hydrochloric acid Sulphuric acid Borax

TABLE

AT RANDFONTEIN ESTATES,

1946

0.175 lb. per ton milled 1.955 0.051 0.015 0.046 0.059 0.003

leaching goes direct to precipitation. Since grinding is in a water circuit, 30 per cent of the precipitated solution, carrying 0.017 dwt. Au and 0.012 per cent KCN, is wasted to maintain solution balance. The precipitate after acid treatment is washed, pressed in Johnson pressure filters, calcined, and charged into No. 100 crucibles with a flux of sand, borax, and manganese dioxide. Crucibles, 30 per charge, remain in the furnace 3 to 4 hr. and, when poured, yield buttons of about 100 oz., which are remelted and poured into 1000-oz. bars for shipment to the Rand refinery. Slag is ground wet, run over riffles to recover possible gold, and shipped as a by-product after assaying for gold, silver, and platinum.

362


Over-all extraction (sand and slime) is 95.4 per cent. Table 63 gives assays for 1946 in dwt. gold pel' ton. Lime is added to mill bins as burned lime and to agitation as milk of lime, about 65 per cent of the total being added to the mill bins. The lead nitrate is added to pregnant solution before clarification. West Rand Consolidated (Type 1155). This company is operating two reduction plants which have a combined rated capacity of 240,000 tons per month. At the north plant, a special pyrite-treatment section produces a high-grade pyrite concentrate which is sold for sulphuric acid manufacture, Table concentrate at 8"1:.8 pel' cent pyrite produced at the south plant is trucked to the north plant pyrite section. The average pyrite content of the ore for 1947 was 2.75 per cent FeS2. The following flow sheets and information, based on 1947 operations, were furnished by the operating staff through F. A. G. Maxwell, consulting metallurgist of General Mining & Finance Corporation. At the north plant, which treats about 60 per cent of the ore, primary crushing is done in Hadfield jaw crushers 30 by 18 in. set at 8 in. and operating 22 hr. pel' day. Primary screening is in trommels with 27'2"-in. holes; secondary crushing is done in Newhouse gyratory crushers set at ~i in. Sorted waste amounts to about 9 pel' cent. Secondary screening is on Robins Gyrex screens with 17'2"-in. openings, and tertiary screening is on Vibrex and Gyrex screens, with %- and %-in. openings. About 40 per cent of the tonnage treated goes to the tube-mill bins and 31 per cent to the stamp bins. The remaining 21 per cent is 76 per cent pebbles, fed proportionately to the 18 tube mills, and 24 per cent washing fines, which are pumped to the cyanide plant. There are 100 stamps serving as tertiary crushers, reducing the minus 17'2"-in. feed through screens with 0.375-in. opening. The crusher-plant washings as undersize of the %-in. screens are dewatered in DOlT classifiers. Rake product goes to the tube-mill bins. Overflow after thickening is pumped to the bowl classifiers in the cyanide plant. There are twelve 67'2"- by 20-ft. tube mills in the tube-mill section, and six following the stamps, five of which are 5 by 16 ft. and the other 6 by 16 ft. Each tube mill is in closed circuit 'with a 6-ft.-diameter HardySmith jig and a DOlT classifier. In the tube-mill section the classifiers are 8 ft. wide, while those in the stamp-mill section are 47'2" ft. wide. Grading of classifier overflow is 48 mesh with 56.8 pel' cent minus 200.


363

The rougher jig concentrate, averaging about 15 dwt. gold and 13 pel' cent pyrite, is cleaned in three 8-ft.-diameter Hardy-Smith jigs, and cleaned concentrate is treated on 12 James tables. Jig and table tails are returned to the tube-mill classifiers. A table middling product is ground in a 2-ft. 10-in. by 4-ft. ball mill, and returned to the tables. Table concentrate goes to the pyrite-treatment plant. There are four DOlT bowl classifiers, 16 ft. in diameter, with 6-ft.-wide rakes, receiving the minus 48-mesh overflows from both the stamp-mill and tube-mill sections. Bowl classifiers make a sand-slime separation at 100 mesh. The bowl classifier sands are treated in four ~econdary Hardy-Smith jigs before going to the sand-leaching plant. The sands are leached in 18 tanks, 6 of which are 42 in. in diameter by 8 ft. deep and 12 are 40 ft. in diameter by 12 ft. deep. The minus 100-mesh slimes, which average about 1.24 dwt. gold. are dewatered in eight 50- by 12-ft. plain slime collectors and six 50- by 12-ft. DOlT thickeners. Thickened slimes are batch-agitated for 12 hr. in six Pachuca tanks 22 in. in diameter by 35-ft. cylinder and 18-ft. 3-in.-deep cone bottoms. Average specific gravity of agitator pulp is 1.42. Each agitator requires about 300 cu.ft. per min. of free air at 35-lb. pressure, The agitator discharge is filtered direct on drum filters, 14 ft. in diameter by 16 ft. long, six of which are Oliver and one Fraser and Chalmers. Cake moisture averages 26 per cent, and with only a barren-solution wash of 1.06 tons solution per dry ton of slime, the average dissolved loss is 0.013 dwt. pel' dry ton filtered. In the south plant the Hadfield crushers am set at 8 in. and are followed by two washing trammels with 4-in. openings. The plus 4 in. goes to sorting belts, where about 6.6 per cent waste is discarded, and thence to two 6-in. pebble-sorting trommels. The undersize, minus 4 in., from the washing trommels is rescreened on two 17~-in. Ty-rock screens, the plus 13/z in. of which joins the undersize of the pebble trommels and goes to a 4;'ri-in. Symons cone crusher. The minus l~~-in. undersize of the Ty-rock screens is rescreened on two Gyrex screens with %-in. openings. The minus % in., which contains all the washings, goes to lO-mesh dewatering screens and a DOlT classifier. Dewatered oversize goes to the tube-mill bins. Classifier overflow is dewatered in a 50-ft.-diameter Don thickener, and the thickened underflow, amounting to about 150 dry tons per day, goes to the sand-slime bowl classifiers in the cyanide plant. Cone-crusher product is screened on a 3-~-in. Gyrex screen. Oversize of all Gyrex screens goes to two Symons 4-ft. short-head crushers set at .H

364


in. and thence to the mill bins, together with the undersize of the conecrusher screens and the oversize from the dewatering screens and the classifier. Grinding is done in four 6- by 8-ft. ball mills and eleven 67~- by 20-ft. tube mills. Each mill is in closed circuit with an 8-ft.-cliameter HardySmith rougher jig and an 8-ft.-wide DOlT classifier. Rougher jig concentrates are retreated on cleaner jigs, and cleaner-jig concentrate is tabled on James tables. Final concentrate from the James tables, amounting to about 46 tons per day, is trucked to the pyrite treatment at the north plant. Cleaner-jig tails and James table tails return to the ball-mill circuit. The closed-circuit classifier overflows, at 48 mesh with 55.8 per cent minus 200 to two 20-ft.-diameter Don bowl classifiers, where a sand-slime separation at 100 mesh is made. Treatment of sand and slime is the same as in the north plant, except that all slime thickening is continuous in ten 50-ft.-diameter DOlT thickeners and that the sands are not treated on jigs before leaching. The data in Tables 65 to 68 are average results for 1947, with 114,630 tons pel' month milled at the north plant and 82,770 tons at the south plant: TABLE

65.

SCREEN GRADINGS OF BOWL CLASSIFIERS AT WEST RAND CONSOLIDATED

North plant Mesh

Feed, per cent

South plant Slimes,

Sands, per cent

per cent

Feed, per cent

Sands, per cent

Slimes, per cent

1.7 48.4 42.2 7.7

.... 2.8 20.2 77.0

1.7 35.3 35.8 27.2

2.4 47.5 41.5 8.6

3.8 21.0 75.2

i

+48 +100 +200 -200

I

I

1.2 35.5 36.0 27.3 TABLE

66.

STRENGTH OF SOLUTIONS

North plant Item

Solution entering agitators ......... Precipitated solution .............. Strong leach solution .............. Weak leach solution ............... Grinding circuit ...................

South plant

NaCN,

CaO,

NaCN,

CaO,

per cent

per cent

per cent

per cent

.---

---

0.020 0.018 0.020 0.024 0.020

0.0155 0.018 0.052 0.012 0.009

0.016 0.018 0.060 0.008 0.008

0.020 0.018 0.019 0.012 0.020


67.

ASSAY VALUE OF SOLUTIONS

Solution

Pregnant solution, dwt Barren solution, dwt Strong leach solution, dwt Weak leach solution, dwt Tons solution precipitated TABLE

365

68.

:

. . . . .

North plant

South plant

2.68 0.017 1. 76 1. 76 4400

2.045 .0.0l7' 1. 75 0.50 2800

CHEMICAL CONSUMPTION oJ

Chemical

Cyanide (NaCN), lb. per ton Lime (CaO), lb. per ton Zinc dust, lb. per ton

. . .

North plant

South plant

0.265 2.280 0.066

0.266 2.480 0.053

In the combined plants 27.4 pel' cent of the tonnage milled is treated by sand leaching and 71 per cent by slime agitation; 1.6 per cent is produced as concentrate for treatment in the pyrite plant. Combined over-all extraction (sand, slime, pyrite) for both plants was 96.5 per cent on an average head value of 3.536 dwt, gold. The washed residues were Sand Slimes Pyrite

"

0.170 dwt. gold per ton 0.075 0.794

Dissolved gold loss in sand residue was 0.015 dwt., in slime residue 0.013 dwt., and in pyrite residue 0.010 dwt. gold per ton. Tailings dam solution amounted to 26,616 tons per month and assaying 0.016 dwt. was returned, which, when credited to the slime filters, reduced dissolved loss in slime residue to 0.01 dwt. At both plants the Merrill-Crowe precipitation system is used. The gold slimes, after sulphuric acid treatment, are calcined and smelted with sand, borax, Mn02, and sodium carbonate. Other chemicals consumed in pounds per ton are Lead nitrate added to the pregnant solution Chloride of lime added at the crushers Sulphuric acid for gold slimes treatment

0.019 0.026 0.067

Pyrite Plant. In the pyrite-treatment plant the James table concentrate from both the north and the south plants is treated on three primary corduroy tables and ten cleaner tables. Corduroy concentrates go to the north plant amalgam room for treatment. Corduroy-table

366


pyrite tails are ground in two 5;!/z by 8~~-ft. ball mills, each operating in closed circuit with an 8-ft.-diameter Dorr bowl classifier. The reground pyrite at 85.8 pel' cent minus 326 mesh is dewatered in a 30-ft.-diameter Don thickener to 1.23 specific gravity and goes to intermittent agitation in seven Crosse tanks 16 ft. diameter by 11 ft. cylinder and 22-ft.-deep cones. Overflow of the thickener is clarified in three 30-ft.diameter sand clarifiers and sent to the north plant precipitation room. Each Crosse tank holds a charge of 40 dry tons. Average filling time is 10 hr. After filling the pulp is circulated by pump for 1 hr. during which 180 lb. of aorocyanide is added. The charge is agitated with air for 10 hr., then circulated by pump with barren solution added, and continuously decanted until treatment is completed. The decanted pulp at 2.42 specific gravity is discharged to two 25~ft. diameter by 8-ft.-deep surge tanks and filtered on one 8-ft.-diameter by 12-ft. drum filter. First decant solution joins the thickener overflow going to clarification, while the lower grade decant solution and the filtrate are sent back to the north plant slime collectors. The filter cake containing 12 per cent moisture is dried in two steamjacketed screw conveyors to 6 per cent moisture and loaded into trucks for shipment, Total pyrite concentrate produced in 1947 from 2,369,000 tons of ore milled amounted to 38,100 tons containing 82.0 per cent FeS2 and assaying 0.794 clwt. gold per ton. ALL-SLIME PLANTS INTERMITTENT TREATMENT

New State Areas Mill (Type IIa). The New State Areas mill in the Witwatersrand area of South Africa treats 4900 tons of mine ore per clay. The information following has been supplied by the operating staff through Fred Wartenweiler, consulting metallurgist. The ore is of the character usually found in this area and is described as quartz banket reef. It assays 3.8 dwt. per ton gold and 0.35 dwt. per ton silver and contains some pyrite. Crushing. The mine ore is screened and washed on vibrating screens, operating in parallel. The Ty-rock screens are clouble deck with a 2-in. protective screen on the top deck and 17:4 in. square mesh screen on the lower decks. The combined oversize (plus 17-§ in.) is sorted on four 36-in. sorting belts, where about 20 per cent, or 1000 tons, per day is discarded as waste and about 145 tons of primary pebbles are sorted out. The undersize

TREATMENT

OF GOLD ORES

367

Spray water Double deckscreens tap deck 2'; bottom deckI f/'

FIG. 83. Flow sheet of the New State Areas mill, Witwatersrand, South Africa.

368


(minus U~ in.) is rescreened and dewatered on a 10- by 4-ft. double-deck Ty-rock screen. The sorted ore is crushed in two 30- by 18-in. Hadfield jaw crushers and one 3-ft. Symons standard cone crusher. Crusher product and the oversize (plus ~i in.) of the 10- by 4-ft. dewatering screen go to a 12- by 5-ft. double-deck Ty-rock screen fitted with 2-in. upper deck and %-in. lower deck screens. Washing fines as undersize of the dewatering screen are dewatered in a DOlT classifier followed by a cone and a settling tank. Classifier rake product goes to the mill bins, and overflow after thickening is pumped to the secondary grinding circuit. Final crushing of the tertiary 12- by 5-ft. Ty-rock screen oversize is done in two 5H-ft. Symons short-head crushers, the product from which joins the screen undersize and goes to the mill bins. From the top deck of the tertiary screen the plus 2-in. oversize is diverted as required to the secondary pebble bin. About 90 tons per day of secondary pebbles is used. Crushing and sorting cost for 1947 was 0.65s (13 cents) per ton milled. The crusher-plant product to mill bins averages 5.0 per cent plus 1 in. and 49.0 per cent minus 7~ in. Grinding. Primary grinding is done in nine 6~- by 20-ft. tube mills, each in closed circuit with a DOlT duplex classifier. Eight classifiers are 8 ft. wide, and one is 7 ft. wide. The primary-overflow sizing is 6.0 per cent plus 48 mesh, 38.0 per cent plus 100 mesh, and 38.0 per cent minus 200 mesh. Tube-mill discharge flows over nine corduroy tables, each 5 ft 10 in. wide. The corduroy-table concentrate is cleaned on one Wilfley and one James table and then amalgamated. Corduroy-table recovery is about 24 per cent of the total gold recovered. The primary-classifier overflow passes over nine more corduroy tables and then goes to intermediate classifiers, three of which are DOlT bowl classifiers with 20-ft.-diameter bowls and three are J.C.L. spiral-pan classifiers with 16-ft.-diameter pans. The rake product amounting to about 2400 per day goes to five 6~- by 22-ft. secondary tube mills, each in closed circuit with a 7-ft.-wide DOlT classifier, and one 57~- by 20-ft. open-circuit tube mill. Each closed-circuit tube-mill discharge flows over a corduroy table similar to those in the primary circuit. The overflow of the secondary classifiers returns to the intermediate classifiers. Grinding is in a cyanide circuit. Slime Treatment. Overflow of the intermediate classifiers is thickened to 45 per cent solids in nine 76-ft. square intermittent settling tanks with 45-deg. sloping sides and 2X in. per ft. bottom slope.


369

The thickened slime is pumped to fomteen 21-ft.-dianieter by 45-ft.. Brown tanks, where it is agitated for 18 hr. at 38.5 per cent solids. Dilution of thickened slime is with Butters filtrate used in hosing and.discharging the 76-ft. square slime collectors. Filtration is done on 600 Butters leaves 9 by 4.5 ft. and' two 16-ft.~, diameter by 14-ft. Fraser and Chalmers filters. Part of the Butters filtrate is used for hosing down slime collectors. The balance, together with all rotary filter filtrate, goes to precipitation. About 1.2 tons barren wash ,is used per dry ton filtered. Dissolved loss in tailings is given as 0.015 dwt. and undissolved 0.18 dwt. per dry tsm. Pregnant solution is clarified in five 40-ft.-diameter sand clarifiers, and approximately 5000 tons per day, averaging 2.05 dwt. per ton, is precipitated by the Merrill-Crows process. Zinc consumption is 0.073 lb. per ton of ore. Solution strength is 0.34 lb. N aCN and 0.30 lb. CaO per ton, and consumption is 0.45 lb. NaCN and 2.5 lb. CaO per ton milled. Lead nitrate is added to the clarifiers ahead of the Crowe tank, Refining. The precipitate is acid-treated with 12 per cent sulphuric acid. The residue is water washed, filter pressed, and then calcined before smelting. Sand, borax, fluorspar, soda ash, and a small amount of manganese dioxide are used for smelting flux. The buttons are remelted into 1000-oz. bars. The amalgam is retorted and remelted with the buttons. Tails from the amalgam barrel are sold as a by-product. West Springs Mill (Type IIa). The West Springs mill in South Africa receives about 4000 tons per day of which approximately 340 tons is discarded by sorting. Crushing is done by two Gates gyratory crushers and two 7-in. Newhouse gyratory crushers. The crushed ore to the tube-mill bins averages about 90 per cent minus 1 in. Primary grinding is done in eight 6>~ by 20 tube mills, operating in closed circuit with eight 8- by 18-ft. 4-in. Dorr duplex classifiers. A circulating load of about 1.2 to 1 is carried. The mill discharge is screened on 1H and H in. The plus U~ in. is used as secondary pebbles. The minus U~ in. plus H in. is crushed in three Hecla disk crushers and returned to the mill bins, and the minus >~ in. returns to the primary classifiers. The primary-classifier overflow at 24 per cent plus 48 mesh and 35 per cent minus 200 mesh goes to two Dorr bowl classifiers 8 by 30 ft. with 20-ft.-diameter bowls operating in closed circuit with six 6>~- by 20-ft. secondary pebble mills. These classifiers rake about 2600 tons of sand per day, and since their installation, improved operation through selective grinding of sulphides has been noted.

370


The secondary mills are loaded with a mixture of 2-in. cast-iron balls and the plus I>~-in. pebbles rejected from the primary mills. Bowl overflow is 9.5 per cent plus 100 mesh and 65.8 per cent minus 200 mesh. The pulp is thickened in seven 70- by 12-ft. plain collecting tanks and is discharged at 30 per cent moisture. Agitation after thickening is done at 52 per cent solids in twelve 15- by 45-ft. Pachuca agitators. Total period of agitation is 8H hr. From 850 to 900 cu. ft. per min. free air is supplied to the agitators at 35 lb. per sq. in. pressure. Filtration is on 490 Butters leaves, and barren wash amounts to about 1 ton per dry ton filtered. Tailings losses are given as 0.02 dwt. soluble and 0.17 dwt. undissolved gold per ton. Zinc dust is used to precipitate 3800 tons of solution per day assaying 3.5 dwt. gold per ton. Solution strength in the cyanide plant is 0.018 per cent in both KCN and CaO, and the consumption amounts to 0.28 lb. KCN and 1.75 lb. CaO per ton milled. Cyanide strength in the grinding circuit is 0.008 pel' cent. Five pounds lead nitrate is added to each agitator, and 2 lb. per shift ahead of precipitation. CONTINUOUS TREATMENT

East Geduld Mines, Ltd. (Type IIa). This plant receives approximately 6400 tons of crude ore per day, of which 6H per cent is rejected as waste. After screening, the ore is crushed in two Hadfield jaw crushers before washing and sorting. The product of primary crushing is screened on 4, 1H and %-in. grizzlies and screens, and sorting is done on three sizes of oversize products. The sorted ore is further reduced in two 4:!/z-ft. standard Symons and two 4-ft. short-head Symons cone crushers. The crushed ore and sands of the washing fines classifier go to a 5000-ton tube mill bin. Tube-mill feed is all %-in. screen undersize, which averages about 39 per cent plus % in. and assays 5.68 dwt. gold, 0.57 dwt. silver, and 3 per cent pyrite. Primary grinding is done in five 8- by 16-ft. tube mills operating as opencircuit mills except for a portion of bowl classifier sand, by the return of which about 22 per cent greater original feed tonnage can be handled. Secondary grinding is done in fourteen 6:!/z- by 20-ft. tube mills each in closed circuit with an 8-ft.-wide DOlT classifier and the tertiary .in six 8- by 16-ft. tube mills in closed circuit with six 18-ft.-diameter bowl classifiers.


371

Primary pebbles, plus 4 in., average 14 lb. each. Secondary pebbles, minus 4 in. plus 17~ in., average 3 lb., and tertiary pebbles, minus 1~ in. plus IX in., average %: lb. each. Primary and secondary pebbles are hand-sorted. Tertiary pebbles are the oversize of Ui-in. grizzlies. A composite grinding load of 5 tons steel and 11 tons pebbles is carried in the secondary mills. The primary and tertiary mills carry pebbles only. Grinding is in an alkaline-water circuit, and the ground product averages SO per cent minus 200 mesh. The slime is thickened to 57 per cent solids in sixteen 50-ft.-diameter two-compartment Dorr tray thickeners and three 70-ft.-diameter singlecompartment Dorr thickeners. The thickened pulp, after diluting to 55~

FIG. 84. Aerial view of East Geduld Mines, Ltd., illustrating characteristic South African tailings piles and out-of-door installations of thickeners and agitators.

per cent solids with filter effluent, is continuously agitated for 35 hr. in thirteen 50-ft.-diameter by IS-ft.-deep agitators, eleven of which are Dorr and two Denver type. Cyanide is added to bring the strength up to 0.03 per cent KCN in the first agitator, and the solution strength in the last agitator is 0.017 per cent KCN. Chemical consumption is 0.46 lb. KCN, 2.21 lb. CaO, and 0.05 lb. Pb(NO a)2, which is added to the agitators and to pregnant solution before clarifying. The agitated pulp is filtered and washed on twelve 14-ft.-diameter by 16-ft.-Iong Oliver and two Fraser and Chalmers filters of the same size. Barren-solution wash, amounting to 0.9 ton per ton of solids, is used. Cake is discharged at 26 per cent moisture and is weighed on a Merrick

372


4 "grizzlies

2-Sortingbelts I~"grizzley ~" Tyrock screens

+oE,..--C.:::.:~"'us"'h""i!X1.l2lantfines

r-

Borren solution

FIG. 85. Flow sheet of the East Geduld mill, Witwatersrand, South Africa.


373

weightometer, repulped with 0.4 ton of barren wastage and water, and pumped to the slimes dam. Undissolved gold in residues averages 0.20 dwt. per ton. Precipitation of the gold and silver from 1.28 tons of solution per dry ton treated is by Merrill-Crows precipitation, using 0.064 lb. zinc per dry torr. Over-all extraction is 96.5 per cent. Total treatment cost for 1946 was 36d (60 cents), of which cyaniding was 1l.7d (19.5 cents). Total power consumption was 26.4 kw.-hr. per ton milled, of which the crushing plant took 2.6 kw.-hr., grinding 20.2 kw.-hr., and cyaniding 3.6 kw.-hr. Sub-Nigel Gold Mining Co., Ltd. (Type Ha), This mine is the property of New Consolidated Goldfields and is located in the extreme southeastern area of the large tonnage operations of the East Rand. The following information was supplied by the operating staff through A. Clemes, consulting metallurgist. The reduction plant is an all-slime plant of 2500 tons' daily capacity, crushing and grinding take place 6 days per week, and cyaniding 7 days. The ore averages 9 dwt. gold and 0.9 dwt. silver per ton and contains about 2.7 per cent pyrite and 1.0 per cent pyrrhotite. The are from the mine bins is screened on a 3-in. grizzly with minus 3-in. undersize rescreened on two double-deck Allis-Chalmers low-head screens (17'2- and %-in.) before washing. The plus 3 in. is washed on the lower end of the sorting belt. The minus 3 in; plus 17'2 in. is washed on short, inclined washing belts, and the minus 17'2 in. plus % in. is washed on a %-in. washing screen.. The washing fines from belts and screens are dewatered in two 6-ft. Dorr classifiers and one 50-ft.-diameter Dorr thickener, with classifier rake product going to the mill bins and thickened classifier overflow pumped to the secondary grinding circuit. Sorting is done on two sizes, plus 3 in. and minus 3 in. plus 17'2 in. Primary pebbles are taken from the coarse sorting belt, and secondary pebbles from the fine sorting belt. The sorted ore and the oversize from the %-in. washing screen are crushed to % in. in four No. 7 Newhouse crushers. Crusher product is conveyed to the mill bins. Grinding is done in two stages, using both ball mills and pebble mills for primary grincling and pebble mills only for secondary grinding. The primary ball mills were installed in 1936 because of insufficient supply of suitable-size pebbles for primary grinding. Composite loads of balls and pebbles were tried but were found unsuitable because of the poor quality of coarse pebbles. The grinding is in a water circuit. The primary ball mills are 6 ft. 6 in. diameter by 9 ft., and the primary pebble mills are 6 ft. 6 in. by 20 ft. Secondary mills are 6 ft. diameter

374


I Mineore I 'f'

+3/1

3/1Grizzley

t

10. ,'-,- -----"--'

Double deck screens I Vi' a :}4'

I

+

-'+:wt

'2

Fines

r---I Washing screen

t,

2-Washing belts 12-SorIing belts

...

I

Waste bin

OncbrYI Debbles I I ISeC

I crushers INewhouse

Sands

_'f'

Classifier

I

".

I

'Mil/bins y

I

I

Primary mil/s

I

I

Corduroy tables

...

Thickener I

I Retum water L-..::=:'-...J

I-...j Primary classifiers I y

I Bowlclassifiers Tubemi//s I IAmalgam ~ room Corduroy tables

,k

I

I

Return water '--....:...;.;;-'----' fYJ!!lide

I

3-Bro:n tanks

I

9-Oorragitators IO-Oliver filters

ISand clarifiers

...

I

Mem'll-Crowe precipitation

.....--_---''f'..,

I SmeltinQ I

I

I

I solution Barren I r-..I-:='f'_ _..,

.

I L...,..J Barren solution

01

0'"

I

r-J

" /I-Darrthickeners

I

Barren solution

IWater andBarren • solution I Repu(oersl

...

I

Slimes dam

I

FIG. 86. Flow sheet of the Sub-Nigel mill, Witwatersrand, South Africa.


375

by 20 ft. long. Primary grinding is to about 35 mesh with 50 per cent minus 200, and secondary grinding to about 75 per cent minus '200. Corduroy tables are used in both the primary and secondary grinding circuits. Fifty-six to fifty-seven per cent of the total gold is recovered on the corduroy tables. The corduroy concentrate is cleaned on a Wilfley table, with table tails returned to the grinding circuit. Table concentrate is amalgamated and retorted after amalgam cleaning on two 4-ft. underdriven bateas, The overflow from the bowl classifiers gravitates to eleven 50-ft.-diameter DOlT thickeners, three of which are two-compartment open-type trays. The thickener is controlled by diaphragm pumps at about 63 per cent solids and flows to a tonnage-measuring box where constant check on mill tonnage is made. From the tonnage box the pulp is pumped to three Brown tanks for pre-aeration under close pH control. The period of preaeration is 12 hr., and alkalinity is maintained at a pH of 9.6. Adoption of pre-aeration followed extensive investigation and testing to determine the cause of intermittent periods of high residues and high cyanide consumption. An account of these investigations is given in a paper by King, Clemes, and Cross" presented at the February, 1947, meeting of the Chemical Metallurgical and Mining Society of South Africa, Mter aeration the pulp is diluted with barren solution to about 40 per cent solids, the cyanide is added, and the pulp is continuously agitated for approximately 42 hr. in nine 50-ft.-diameter by 16-ft. Dorr agitators. Cyanide and lime strengths at the beginning of agitation are 0.02 per cent KCN and 0.002 per cent CaO. Lime is added to the last agitator to a solution strength of 0.01 per cent CaO, and the pulp is then filtered and washed on ten Oliver filters, five of which are 11 ft. 6 in. diameter by 14 ft. and five are 14 ft. diameter by 16 ft. Filter cake at 27 per cent moisture is repulped and pumped to the slimes dam. Return water from the dam is pumped back to' the mill solution tanks. Oliver filtrate is clarified in five sand clarifiers, two of 40 ft. diameter and three of 30 ft. diameter with sand beds 18 to 24 in. deep. Sand used for clarification is the rake product of the primary classifiers. After clarification the rich solution goes to Merrill-Crewe precipitation. Zinc-dust consumption is 0.1 lb. per ton of ore. Precipitation ratio is 1.7: 1. Rich solution is about 2 dwt. per ton, and barren solution 0.02 dwt, Over-all recovery is 96.8 per cent. Lead nitrate is added prior to filtration. The gold mud from Merrill-Crewe precipitation is acid treated, calcined, 23 "The Treatment of Gold Ore Containing Pyrrhotite at the Sub-Nigel, Ltd.," Jour. C.M. and M.S.S.A., Vol. 47, No.8.

376


and smelted. Gold buttons are remelted, and 900-oz. gold bricks poured, which are shipped to the Central Rand refinery. Van Dyk Consolidated Mines Ltd., (Type IIa). This plant, with a milling capacity of approximately 4000 tons per day, is one of the newer plants of the Union Corporation. The ore from the mine bins, with average maximum size of 7 in. and occasional pieces up to 14 in., passes over a grizzly tapered from 27:! to 3 in. Grizzly oversize is washed in chutes and passes onto the coarse sorting belt. The undersize is washed and screened on a 4- by 8-ft. Ty-rock screen with 2:>:1 in. square mesh opening. The washed oversize from the Ty-rock screen goes to the intermediate sorting belt. The undersize is rescreened on three 4- by 8-ft. double-deck Aero-vibe screens, the top screens of which are I-in. square mesh protective screens. The bottom decks are fitted with 7~-in. square mesh for the first 2 ft. to remove water and fines of the washing operation. The remaining 6 ft. is %-in. square mesh screen. The oversize from both decks passes onto the fine sorting belt. The three sorting belts are 36 in. wide and run at a speed of 75 ft. per min. Primary tube-mill pebbles as well as waste are picked from the coarse sorting belt, while secondary pebbles are diverted from the discharge of the intermediate sorting belt to the pebble bin, as required. The average amount of waste sorted from 1,341,000 tons of crude ore in 1947 was 14 per cent, or approx..rimately 600 tons per day of 16 hr. The sorted ore from the coarse sorting belt is crushed in a 30- by 23-in. Hadfield jaw crusher set at 4 in. and thence joins the sorted ore from both the intermediate and the fine sorting belts going to the secondary crusher bins. The ore is then further reduced in two 4X-ft. (one spare) standard Symons and two 4-ft. (one spare) short-head Symons cone crushers. The shod heads are in closed circuit with three 4- by 8-ft. Aero-vibe doubledeck screens with %-in. J-in. square mesh protective top screens and %-in. square mesh bottom screens. The minus .%-in. screened product goes to the tube-mill bins together with the undersize from the secondary Aero-vibe screens and the classifier rake product of the washing fines. Overflow of the classifier is dewatered to 35 per cent solids in a DOlT thickener, the spigot of which is pumped to the bowl classifiers of the secondary grinding circuit. In the grinding department there are four sections, each having one 8by 16-ft. primary mill in closed circuit with an 8-ft. DOlT classifier and two 8- by' 16-ft. secondary mills in closed circuit with one classifier. In three of the sections the secondary classifiers are DOlT bowl classifiers with 16-ft.wide model F rakes and 18-ft.-diameter bowls. The fourth secondary classifier is a 5L!-in. Duplex submerged-weir Akins classifier.


377

Both the primary and secondary mills are grate-discharge types driven by 300-hp. motors connected through gear reducers to the center of the mills at the discharge end. Composite loads of pebbles and balls are carried in the milk In the primary mills 4-in. cast-steel balls are used, and in the secondary mills 2-in. white-iron balls are used. The primary mills carry about 7 tons. of balls and 12 tons of pebbles, and the secondary mills approximately
378


tated is 1.4 tons per ton of ore treated. Gold slimes are acid-treated, calcined, and smelted in the usual manner. Table 69 gives chemical consumption, power, and cost distribution for the year 1947. , Marievale Consolidated Mines, Ltd. (Type IIa). This is the newest plant of the Union Corporation located on the East Rand. It has a capacity of 2100 tons per day. The mine ore is coarse crushed, before washing, in two 30- by 23-in. Hadfield jaw crushers. The crushed ore minus 3}-~ in. passes to two double-deck washing screens. These are 4- by 12-ft. Symons. The top deck is fitted with 1}-4:- by l%-in. screen, and the lower deck with IH2- by l}--§-in. screen. The oversize of the top deck is sized on a 2%-in. grizzly for coarse and fine sorting on 30-in. sorting belts running 80 ft. per min. The oversize of the bottom deck minus 1% in. plus 1 H2 in. goes to Symons short-head crushers. The undersize from the first 6 ft. of the lower deck is washing fines. These are dewatered in an 8-ft. Don classifier, the TABLE

69.

CHEMICALS IN POUNDS USED PER DRY TON MILLED AT THE

V AN

Cyanide as (NaCN) Lime (100% CaD) Lead nitrate

. . .

0.394 1.896

0.032

DYKE MILL

Sulphuric acid Hydrochloric acid Zinc Borax

. . . .

0.041 0.056 0.073 0.004

rake product of which joins the balance of the minus IH2-in. undersize going to the mill bins. The overflow after dewatering in a 70-ft. Dorr thickener is pumped to the secondary mill classifiers. The washing fines amount to about 470 tons per day, of which 25 to 30 per cent is minus 80-mesh classifier overflow by-passing grinding. The sorting belts carry about 70 tons per hr. from which waste is sorted by 38 pickers (22 on coarse and 16 on fine). The amount of waste in 1947 was 12.8 per cent of the crude ore. Ore from the fine sorting belt is split on a l%-in. grizzly. The undersize is used for pebbles in the secondary grinding mills, and the oversize together with the ore from the coarse sorting belt goes to a 474:-ft. Symons standard cone crusher and joins the product from the two 4-ft. Symons short-head crushers, going to the final screening on one 4- by 10-ft. Symons and two 4- by 8-ft. Rod-deck screens. These screens are in closed circuit with the short-head crushers, so the final crushed product is screen undersize of IH2- by 17-&-in. slotted screen. Coarse sorting ranges from 2H to about 8 in. maximum size, and fine sorting ranges from 1}-~ to 2~ in.


379

Iff!!, Coo andPbN

FIG. 87. Flow sheet of the Marievale Consolidated mill, Witwatersrand, South Africa.

380


Burned lime amounting to 0.8 lb. per ton milled is added at the primary crushers. The crushing plant operates 10 hr. per day 6 days per week. Milling. There are two grinding sections, with primary ball mills and secondary composite mills, each section handling about 970 tons per day. Each section has one 8- by 8-ft. primary ball mill in direct closed circuit with one 72-in. simplex Hi-weir Akins classifier and two 8- by 16-ft. secondary mills in closed circuit with one 54-in. duplex submerged Akins classifier. The primary mills have %- to l-in. tapered discharge grates, operate at 20 r.p.m., and are driven by 350-hp. motors through gear reducers with slow-speed shaft direct-connected to the mill at the discharge end. Power input is 281 hp, Ball charge is approximately 25 tons, using 54 per cent 4-in., 33 per cent 3~-in., and 13 per cent 3-in. balls for daily ball additions. Average mill density is 80 per cent solids, and ball consumption is 1.9 lb. per ton milled. Primary classifiers operate at 5 r.p.m. and are fitted with special tooth-edged wearing shoes that prevent building up under the flights and have improved the life of wearing shoes. Primary overflow at about 50 per cent solids is minus 14 mesh with 30 per cent minus 200 mesh. Original feed is controlled by vibrating pan feeders, and feeder belts are equipped with tonnage indicators. This is one of the few plants on the Rand where feed scoops are used on mills to complete the closed circuit instead of pumping the mill discharge to the classifiers. The secondary mills have ~~- to %-in. taper round-hole discharge grates. They are also driven by 350-hp. motors direct-connected to the mill through speed reducers. One mill of each section operates at 24.5 r.p.m., and the other at 19.5 r.p.m. The average power for the two mills is 296 hp., the faster mill pulling about 52 per cent of the total power. The secondary mills carry a composite charge of about 12 tons made up of ore pebbles, clrill-steel slugs, and2-in. white-iron balls. Each mill uses about 25 tons of pebbles and 600 lb. of slugs and balls per day. Secondary mill discharge is held at 74 per cent solids, and slug and ball consumption is 1.1 lb. per ton milled. The ground product overflowing secondary classifiers is 78.5 per cent minus 200 mesh and flows by gravity to five 70-ft.-diameter DOlT thickeners. Thickened discharge at 63.7 per cent solids is controlled by Dorrco diaphragm pumps. Grinding operation is 6 days per week and is in water of 0.008 per cent NaOH alkalinity. Cyaniding. The thickener discharge is diluted to 1.2 to 1 (45.5 per cent


381

solids) with filter effluent and goes to continuous agitation in five 50-ft.diameter by 20-ft.-deep DOlT agitators. Cyanide, lead nitrate, 'and milk of lime are added. The agitation period is 40 hr. Agitation ail' supplied from the mine compressors is reduced to 20 lb. pel' sq. in. 'pressure, and the amount used is approximately 1900 cu. ft. per min. of free 'air. . The agitated pulp is filtered on four 14-ft.-diameter by 16-ft.-face Fraser and Chalmers drum filters. Filters have variable-speed drive with a speed range from 2X to 3H min. per revolution. Washing is with precipitated solution assaying 0.01 dwt. gold, and the amount of wash used is 0.9 tons per ton of solids filtered. ., 70.

TABLE

GOLD VALUES AT MARIEVALE

Crude ore

5.25 dwt. per ton 0.25 Mill hea.ds .. . . . . . . . .. 5.98 VVashed residues, , . .. 0.22 Pregnant solution 4.50 Barren solution 0.01 VV~ste

TABLE

71.

CHEMICAL CONSUMPTION AT MARIE VALE

Cyanide (N~CN) .... Lime (C~O). . . . . . . .. Lead nitrate Zinc Muriatic acid Sulphuric acid Borax, The total reduction cost in 1947 was 46.26d (77.1 cent) .per dry ton treated

0.400 lb. per ton milled 1. 500 0.030 0.065 0.030 0.055 0.006

Filter cake discharged at 25.5 per cent moisture is repulped to 53 per cent solids and pumped to the slimes dam. The barren solution necessarily wasted to balance plant solutions is used in repulping filter cake and amounts to 0.23 ton per dry ton of filter. Total loss of value through this source is 0.0023 dwt. gold and 0.092 lb. cyanide per dry ton. Merrill-Crpwe precipitation is used. The gold slimes are acid-treated, calcined, and smelted in reverbatory furnaces. Tables 70 and 71 are average figures for 1947 during which 617,000 tons were milled. Venterspost Gold Mining Company Limited (Type IIa). The treatment scheme followed in this 4400-ton-a-day plant, which is one of the newest of the New Consolidated Goldfield group, is as follows: Crushing Plant. TIllS plant operates 14 hr. per day, 6 days per week.

382


The mine ore is conveyed on a 36-in. belt to one 6-in. grizzly, the oversize from which is crushed in one of two 23- by 30-in. Hadfield jaw crushers. The grizzly undersize (6 in.) flows by gravity to one X-in. washing grid. The washed oversize joins the crusher product and is conveyed on a 3G-in. belt to two 3-iri. secondary grizzlies. The minus :J,~-in. undersize from washing grid is pumped by one of two 5-in. grit pumps to two 8-ft. Dorr F classifiers. The oversize of the secondary grizzlies (minus 6 in. plus 3 in.) is split to two 36-in. coarse-ore sorting belts running 60 ft. per min., while the secondary grizzly undersize (minus 3 in.) goes by gravity to two 4- by 10-ft. Ty-rock vibrating screens with a 2:J,:l:-in. square mesh opening. The oversize (minus 3 in. plus 2X in.) of these screens gravitates to two 42in. fine-ore sorting belts running 60 ft. per min., and the undersize (minus 2X in.) from the Ty-rock screens is conveyed on a 30-in. belt to four 4by 12-ft. Allis-Chalmers horizontal vibrating screens with a %-in. square mesh opening. Waste amounting to 3 per cent of crude ore and averaging 0.30 dwt. gold per ton is picked from both coarse and fine sorting belts. The coarse sorting belts carry about 75 tons per hr. each; the fine sorting belts about 45 tons per hr. Sorted ore passes to three (one spare) 4H-ft. Symons standard cone crushers, making a product (minus 1;!/z in.) which is conveyed on a 30-in. belt (joining the undersize of the Ty-rock screens) to Allis-Chalmers screens. The undersize from these screens passes directly to the mill bins, while the oversize (plus % in.) is crushed first in a 4-ft. Symons short-head crusher. Sands from the crusher-plant classifiers discharges onto a 42-in. belt, which also receives the undersize of the Allis-Chalmers screens and the product (minus % in.) of the short-head crushers. It conveys the product to the mill bins over a Blake-Dennison weightometer. Slime overflow of the crusher-plant classifiers gravitates to two 50-ft. Dorr thickeners. Thickener discharge at 44 per cent solids is pumped to the cyanide-plant thickeners. Grinding Plant. Grinding is carried out in a water circuit in four sections, which receive about 4400 tons per day of new feed including pebbles, the average screen analysis of which is 4 per cent plus % in., 28 per cent plus ;!/z in., 33 per cent plus X in., 4 per cent minus 200 (individual basis). Each one of the four primary grinding sections includes a 9- by 10-ft. ball mill (grate discharge) driven by a 450-hp. motor. Mill speed 17.5 r.p.m.; grate opening is ~-in. Liners are manganese steel of the wave type; life is 160 days. The ball load is 32 tons, 4-in. semisteel cast. Mill input power is 395 hp. The ball consumption amounts to 2.7 lb. per ton milled. Mill discharge is 78 per cent solids. The ball mill is closedcircuited with a 10- by 25-ft. DSFX classifier with a rake speed of 19


383

strokes per minute. The overflow carries 33 per cent solids analyzing 32.6 per cent minus 200 mesh. The sand tonnage raked amounts to about 4000 tons. The mill discharge is pumped to the classifier by an 8-in. pump driven by a 90-hp. motor. The primary-classifier overflow flows by gravity to two DSFXB 8- by 31-ft. 8-in. by 16-ft.-diameter secondary bowl classifiers. . Each of the four secondary grinding sections which receives the classifier sands includes two 8- by 16-ft. grate discharge pebble mills driven by 350-hp.motors. The mill speed is 20.5 r.p.m.; Input power 220 hp. The grate opening is :J/z in. ; pebble size, minus 3 in. plus 2Ji in. The liners are wave type and made of cast iron, with a life of 420 days. Pebble load 18.8 tons; consumption, 18 tons per mill per day. Mill discharge is 68 per cent solids. The rake speed of the bowl classifiers is 13 strokes per minute; bowl speed, 3 r.p.m.; overflow, 12}-2' per cent solids at 72 per cent minus 200 mesh. The sand tonnage raked is about 2000 tons for each classifier. Each mill discharges over six corduroy tables. The table tails are pumped to the bowls by one 8-in. pump with a 100-hp. motor. Thickening. The bowl overflow passes by gravity through distributors to twelve 70-ft.-diameter by lO-ft. Bowley-drive Dorr dewatering thickeners. Thickener discharge is controlled at 1.64 specific gravity (62 per cent solids) by twelve 4-in. Triplex Dorrco pumps. Average value of the solids in the discharge is 2.067 dwt. gold. The thickened slime is pumped to the agitators by two (one spare) 7-in. pumps with 150-hp. motors after diluting with filter eflluent. Strong cyanide and milk of lime are added to pump intake. Agitation-Contin1lOus. This operation is carried out in six 33-ft.-diameter 19- by 29-ft. cone Pachuca tanks. Agitation time is 18 hr.; pulp density, 1.404 specific gravity (45.7 per cent solids). The air used amounts to approximately 250 cu. ft. per min. per agitator at 30 lb. pressure. The air is supplied from the mine at 80 lb. pressure. Solution strength in first agitator = 0.016% KCN, 0.014% CaO Solution strength in last agitator = 0.014% KCN, 0.012% CaO Filtering and Residue Disposal. The filtering equipment consists of six 18-ft.-diameter by 16-ft. Fraser and Chalmers rotary filters. The speed is 1.41 min. per revolution; cake moisture, 25 per cent. Filter data include the following: gold value of caking solution, 1.607 dwt.; of barren solution, 0.01 dwt.; of washed residue, 0.155. Dissolved loss is 0.014 dwt. gold per dry ton. Barren wash is 0.71 ton per illy ton. The filter cake is diluted with barren solution and fresh water to 1.2 to 1 and pumped to the residue dam by one 8-in. pump against 75-ft. static head. The life of the filter cloth is 90 days. Return solution from the dam at a value of 0.07 dwt. per ton is pumped to the dewatering thickeners.

384


Clarifuxuion. This step is carried out in four 55- by 8-ft. sand clarifier tanks, using bowl classifier sand for sand beds. These beds are scraped daily, and the scrapings pumped back to the bowl classifiers. The sand is renewed every G to 9 months. Precipitation and Smelting. The equipment consists of four 40-leaf 72. CHEMICAL CONSUMPTION AT VENTERSPOST Cyanide (KCN) 0.224 lb. per ton milled Lime (CaO)* 0.500 Zinc dust.. .. . . .. .. .. 0.052 Sulphuric acid (H 2S04 ) 0.082 Borax. . . . . . . . . . . . . .. 0.002 Lead nitrate (PbN0 3 ) 0.011 * Lime is added to crushed ore and as milk of lime to agitators. TABLE

TABLE 73. POWER CONSUMPTION Total crushing plant including waste disposal........ 2.2 Kw.-hr. per ton milled Total grinding 15.3 Treatment including thickening, agitation, filtration, and tailing disposal. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. 2.2 Precipation, smelting, and concentrate cleanup....... 0.4 Mill return water pumping........................... 0.5 Total .' 20.6 Kw.-hr. per ton milled TABLE

74.

PERCENTAGE DISTRIBUTION OF DIRECT COSTS

Labor Department

Power

Crushing plant .......... Milling .................. Cyaniding ............... Total .................. TABLE

75.

10 31 10 21

Stores White

Native

14 9 17 12

20 5 11 10

56 55 62 57

GENERAL DATA AT VENTERSPOST

Over gold value of crude ore Over gold value of ore milled Over gold value waste sorted out Over gold value washed residue Ratio of precipitation Barren solution wasted Total gold recovery Gold recovered on corduroy tables Grinding Cyaniding

. , . . . . . . . .

4.336 4.461 0.300 0.155

dwt. dwt, dwt. dwt.

1.01:1

o.6 tons

per dry ton 96.2 per cent 55 per cent 6 days per week 7 days


385

Merrill precipitation units, one 30-in. by 30-leaf gold slimes press, two 12by 10-ft. acid-treatment vats, one 12-tray calcining furnace. Lead nitrate is added to the pregnant solution both before and after clarification. Amalgam Cleanup. The corduroy-table concentrate is cleaned on two Fraser and Chalmers tables with table tails returned to grinding circuit, Fraser and Chalmers table concentrates are amalgamated in 3-ft. by 4-ft. 6-in. amalgam barrels, and the amalgam treated on two underdriven bateas and retorted. The batea tails are passed over corduory for recovery of osmiridium, SOUTHERN RHODESIA

This country has been a regular gold producer since the early days of mining in South Africa. Although most of the deposits are small and widely scattered, a few mines of importance have been developed. A total of 3,360,000 tons of ore was milled in 194'1 with the production of 522,735 oz. gold. The following are perhaps the best known of the operating mines: Bustick Mines, Ltd. This is an all-sliming cyanide plant milling about 12,700 tons per month. The plant has one ball mill and two tube mills, and it is reported that 21 per cent of the values are recovered by concentration. Globe and Phoenix Gold Mining Co., Ltd. is a 40-stamp mill now being converted to an all-slime cyanide plant with fine grinding in two tube mills. Concentration is carried out on jigs and corduroy tables. An average of 11,700 tons was milled in June and July, 1948. Phoenix Prince Gold Mining Co., Ltd. This is an all-sliming cyanide plant using corduroy tables, regrind and amalgamation of the corduroy concentrates, and cyanidation of the amalgamation tailings in the main cyanide circuit. Cam and Motor Gold Mining Co., Ltd. In 1945 this plant was reported to be operating 56 Nissen stamps and six tube mills. The monthly capacity was 26,000 tons, and the yield during that year was £558,401 in gold from the milling of 300,000 tons of ore. Wanderers Consolidated Gold Mines, Ltd. Originally a cyanide plant employing vanner tables and corduroy blanket concentration, the mill was revamped in 1936 to all-flotation. Since practically all the gold is associated with the pyrite, the flotation tailings are discarded. The concentrate, averaging about 6 per cent of the tonnaged milled and assaying 1.5 oz. per ton of gold, is reground to 200 mesh and cyanided. In 1947 a total of 392,000 tons of ore, having an average value of 2.13 dwt. per ton, was milled at a cost of $1.07 per ton.

386


Resende Mines, Ltd. This company operates an all-sliming cyanide plant which contains 40 stamps and two tube mills. About 24 per cent of the gold output is recovered by concentration methods. In 1948 an average of 9000 tons per month was milled with a yield of 1562 oz. gold. GOLD COAST COLONY-BRITISH WEST AFRICA

The gold mines operating OIl the Gold Coast may be classed under two main headings according to the type of ore treated: (1) quartz reefs carrying small percentages of sulphides and graphite and (2) banket, or conglomerate, series. The main strike of the reef runs approximately 20 deg. east of north and extends from a point close to the coast at Takoradi for a distance of about 200 miles inland where it disappears under more recent sedimentaries. There are also more or less isolated outcroppings at points 60 to 100 miles on either side of the main reef, while the banket series parallels it to the southeast. The principal operating mines together with rated tonnage capacity of the mills and types of flow sheet are shown in Table 76. The treatment of Gold Coast ore has always been complicated by the presence of arsenopyrite and graphite and by the fine dissimination of values, though a relatively high proportion of the total gold recovery can in most cases be made on blanket tables, followed by amalgamation of the concentrates. Where serious reprecipitation difficulties due to the presence of graphite have been encountered, roasting has been resorted to, but as in other gold districts, the modern trend is toward the roasting of a flotation concentrate rather than the whole tonnage milled. Stage grinding in ball mills with corduroy strakes handling the mill discharge in each stage and fine grinding of the concentrates is the trend of modern practice, which often avoids the necessity of roasting altogether. At Ariston gold mines, for instance, the original flow sheet included a stamp battery and tube mills, followed by amalgamating pans and tables, filtration, and cyanidation with washing filters. This was modified in 1940 by the substitution of ball mills for tube mills, corduroy' tables, flotation, and cyanidation, while today the stamps have been replaced by Symons cone crushers and the flotation. concentrate is roasted before cyanidation. At Ashanti, the main treatment plant (1940) dried the coarsely ground ore and followed this by dry grinding in Krupp mills, roasting, and cyaniding in leaching vats. The tailings, which still carried 2 to 2~ dwt. gold, were retreated by regrinding in cyanide solution to 75 per cent minus 200 mesh, the free gold removed on corduroy strakes, and the thickened pulp agitated and washed on filters. The new treatment plant, which was


387

built after pilot-plant tests, comprises a modern crushing section, followed by two-stage grinding in ball mills with corduroys in each circuit, amalgamating pans and barrels for handling the concentrate, flotation, and cyanidation of the roasted concentrates. It is an interesting fact that, although the new plant uses the more up-to-date methods of treatment,. the old system did have the advantage that straking was avoided and theft TABLE

76.

PRINCIPAL OPERATING GOLD MINES: GOLD COAST, WEST COAST OF AFRICA

are

Quartz reef Quartz reef Quartz reef Quartz reef Quartz reef Oxidized reef Banket ore Banket ore Banket ore Banket ore

Company

Approx. F* head tons per value, day dwt.j

200 Gold Coast Main Reef Ariston Gold Mines, 700 Ltd. Bibiani Gold Mines, 1200 Ltd. Konongo Gold 400 Mines, Ltd. Ashanti Gold Fields 800 Corp. Marlu Gold Mining 2000 Areas Ashanti Adowsena 332 Gold Fields Gold Coast .Banket 450 Areas Amalgamated Ban- 2000 ket Areas, Ltd. Taquah and Abosso, 800 Ltd.

.,

Treatment

11.0

Cyanidation, blankets, amalgamation. 9.5t Blankets, amalgamation, flotation, roasting, cyanidation. 6.5 Blankets, roasting, § flotation, cyanidation. 9.0 Blankets, amalgamation.flotation, roasting, cyanidation. 23.511 Blankets, amalgamation, flotation, roasting, cyanidation. 3..3 Tremmel washers, ball mills, cyan idation, filtration. 3.3 Single-stage grinding, cyanida tion, filtration. 5.2 As above except C.C.D. washing in thickeners. 3.2 Blankets, amalgamated sand slime sep. and treatment. 5.5 As above, except using amalga mation plates.

* F = rated capacity of mills. A number of these mills are still (in 1948) operating on reduced tonnage. t 1947 data. t Stamps to be replaced by Symons crushers. § Blanket conoentrate tabled and table concentrate roasted before fluxing. II New plant. of the high-grade ore was rendered difficult by the fact that dry grinding coated the ore with graphite. Furthermore, a high recovery (97 per cent) of the gold was obtained, a figure which, in 1940, had not been equaled by the new treatment method. A detailed description of one of the plants operating on each of the general ore types follows.

388

CYANIDATION AND CONCENTRATION OF ORES QUARTZ REEFS

Bibiani Gold Mine (Type IVe). This mine is situated in the western province of the Gold Coast Colony, 63 miles from the nearest railhead at Dunkwa. The author wishes to acknowledge the excellent description of the milling operations at this property appearing in the paper "Treatment at the Bibiani Gold Mine, West Africa" by H. A. McGowan (Bui. 492, I.Jl1. and M. (London), November, 1947), from which the following notes are taken. The modern plant is treating 1200 tons per day of 6.5-dwt. ore by a combination of corduory concentration, flotation, and cyanidation of the flotation concentrates. The following abstracts on the occurrence of gold in the lode are taken from a report by Prof. W. R. Jones, who made an examination of Bibiani ore. The metal occurs (1) in minute fissures of quartz of the sheared type, (2) as small particles-some coarse and some extremely fine (less than 2 microns in diameter)-enveloped in the quartz, (3) as small particles in sulphide minerals of less than 1 micron in diameter, (4) attached to the boundaries of crystals and between pyrite and graphitic material, (5) in microscopic fissures traversing pyrite crystals and grains, (6) in a few cases between pyrite and occasional grains of galena, and (7) as minute grains completely enclosed in pyrite. Test work had made it quite clear that to liberate most of the mineral for flotation it would be necessary to grind the ore to at least 80 per cent minus 200 mesh. The subsequent direct treatment of the concentrate by cyanidation, preceded by grinding to minus 325 mesh in order to liberate the gold enveloped in the mineral grains, completed the circuit. Crushing. Primary crushing is done in two 16-in. McCully gyratory crushers breaking to 2H in., after removal of undersize by grizzlies and Gyrex screen with ~'16-in. openings. Secondary crushing is done' with a 4-ft. standard Symons, and tertiary crushing with a 4-ft. Symons fitted with a fine bowl liner and a 4-ft. short-head Symons. Hummer and Nordberg screens closed-circuited with the fine Symons make a minus ~16-in. product which passes to the mill bins. Grinding. Three 1000-ton fine-ore bins supply the ore to three 8-ft.diameter by 6-ft. long-head Wrightson grate-discharge ball mills by means of three individual 16-in. conveyors, each with its own Merrick weightometer. The ~16-in. ore is ground to approximately 48 per cent minus 200 mesh at a daily rate of 1200 tons. Each mill is in closed circuit with its own corduroy strake table and a 4- by 20-ft. DOlT multizone classifier. The overflow from these primary classifiers is pumped direct to a distributor which feeds two 12-ft. bowl classifiers, while the undersize or rake return is feel back into its own primary mill. I

389


-3-1,000-tont, mJ'l1bins Merrick WeiQhtometers

Reagentfeeders

'f" I 3-8' diam. x 6' long head W(J~htson boll mJ'lls ~--.., 3-Banks of six

rt

e:»19' blanket strakes

3- 4'xI9' DOff multizoneclassifiers

~

t (Overflow)

(Sands)

•

2-12' diam. x 10 '-DOff bowl classifier

~

t

2-8'

(overL.fIOW)

diom(:J~~condarY..

mills

t"

-

3-Bonks of six 2 x 19 blanket strakes Corduroy wrhlng tanks

sufe agitator 4-Banks of 10 Denver flotation cells

Head tank

•

tI

(Cotc';

(Conc)

(Tar,s)

t

Return to ~imary" circuit

+

(Middlings)

t

.

L

Water

Pregnant solution for clarification andprecipJ'lation

t

(71Jilings)

t

t

70 headofcells 100.traction thickeners

4' diam.x 8' bowl classifier closed-circuited with 4' diam. x 8' Ion(/. tube mJ'l1

Roosting furnace fluxing Melting furnace

1

t

2-Concentrating tables

t-

1

70waste ---

50'diam. x 10'thickener

t

8' diam.x 12' Dorr-Oliverfilter

t

3-10'diam. x 10' deep. Wallace agjtators

t

-

~

Filtrate to waste

8' diam. x 12'Doff-Oliver filter-+- Water wash

t

4 -14'x 14' Denveragitators

t

-

8' diam. x 12'DOff-Oliverfilter~ Barren solution washes ;'5~0~'x~/(.!:.O?.:'!!EdeefJDorr thickener

t

L-_--Jf.-.

.

~

Underflow to waste

FIG. 88. Flow sheet of the Bibiani mill, Western Province, Gold Coast Colony.

390


The secondary milling circuit comprises two 8-ft.-diameter by 6-ft. longhead Wrightson mills in closed circuit, each with corduory strakes and its own bowl classifier. The bowl overflow, averaging 70 to 74 per cent minus 200 mesh, is sent at approximately 25 per cent solids to the flotation section for treatment.. The primary mills are loaded with 17 tons of 3-in. chrome-steel balls; the secondary mills with 18 tons of 2- and 2~~-in. balls. Each mill discharge is handled by a 6-in. Wilfley pump which feeds a strake table divided into six runs, 24 in. wide by 19 ft. long. The pulp then gravitates to another 6-in. Wilfley pump and is sent to the classifier in circuit with that particular mill. Each mill has its own strakes operating in closed circuit; over-all extraction by this section averages 52 per cent. The corduroy cloths are washed every 2 hr. into a collecting tank. This is washed out daily into two head tanks feeding two concentrating tables, which produce a clean concentrate averaging 12 oz. bullion per pound avoirdupois. The middlings from each concentrating table join and are pumped back into the primary mill circuit for regrinding and restraking. The concentrate from the two shaking tables is dried, roasted, and smelted. All washing of cloths is done by Africans and is under the close supervision of the European shift operator. Straking is carried out on a pulp containing 50 per cent solids. Trial runs when straking in open circuit on a less dense pulp showed no improvement in the extraction. Flotation. The overflow from the secondary bowl classifiers gravitates to a 30-ft.-diameter by 8-ft.-deep flotation-feed surge tank and thence to four banks of ten No. 24 Denver Sub-A flotation machines, arranged in parallel, via a simple pulp distributor which splits the pulp into four compartments, each feeding its individual bank of cells. The 40 flotation machines treat 1200 tons of ore per 24 hr. Each machine has its own individual drive. The first six cells of each bank yield a finished concentrate, while the last four of each bank produce a middling which is pumped back to the flotation feed surge by two 3-in. Wilfley pumps (one stand-by). All wearing parts-such as impellers and hood wearing plates-are of reinforced rubber and are still in use after 5 years of milling. The finished concentrate from flotation is pumped to a concentrateregrinding circuit, while the tailing gravitates to a 100-ft. Dorr traction thickener, the thickened pulp being sent to the tailing pond. A clearwater overflow is obtained from this dewatering thickener and is pumped to the mill steady-head tanks by two Mather and Platt Plurovanepumps directly driven by individual 30-hp. 440-volt motors. All flotation reagents are fed to the mill circuit.


391

Cyanidation. Approximately 50 tons of flotation concentrate is treated daily in this section of the plant. The concentrate is pumped by two 3-in. Wilfley pumps (one stand-by) to an 8-ft. diameter Dorr bowl classifier, which is in closed circuit with a TABLE

77.

SOLUTION VALUES AT BIBIANI

Pregnant solution from first-stage filtration 20 dwt. per ton Pregnant solution from second-stage filtration... .. . . . . .. . .. . .. . .. 2.5 Pregnant solution from washing thickener........................ 0.05 TABLE

78.

GENERAL METALLURGICAL DATA AT BIBIANI*

., Item

Per cent solids

Primary mill discharge ............... Primary strake discharge ............. Primary classifier overflow ............ Secondary mill discharge ............. Secondary strake discharge ........... Bowl classifier overflow...............

80 55 40 70 50 25

Reagents

Lb. per ton

Sodium cyanide Lime Lead acetate Zinc dust Amyl xanthate Oresylic acid Pine oil

0.34 0.50 0.022 0.02 0.25 0.05 0.05

. 22.5 r.p.m, . 28 strokes per min. . 18 strokes per min. 3.5 r.p.m, . . 3.7 lb. per ton balls 0.90 lb. per ton liners . . 200 per cent . 300 per cent

Mill speeds Primary classifier " Secondary classifier Bowl rakes Steel consumption Steel consumption Circulating load primary mills Circulating load secondary mills Tyler mesh

Flotation feed, per cent

Flotation concentra te, per cent

Cyanide feed, per cent

+100 +200 +325 -325

6.6 25.2 13.4 54.8 100.0

1.4 9.5 8.3 80.8 100.0

7.4 92.6 100.0

* The over-all gold extraction in this mill is 90.8 per cent, of which 55.0 per cent is made on the corduroy strakes. 4-ft.-diameter by 8-ft.-Iong Edgar Allen tube mill fitted with Cromaxsteel liner plates and carrying a load of 2-in. steel balls 50 per cent of the mill volume. The liners of this concentrate regrind mill are of the wedge type and last approximately 12 months before renewal. The bowl overflow averages 93 per cent minus 325 mesh and overflows the bowl lip at 7 per cent solids to a 2-in. Wilfley pump feeding a 50-ft. DOlT thickener. The thickened pulp from the thickener gravitates from

392


a DOlT simplex diaphragm pump to an 8- by 12-ft. Don-Oliver drum filter for final dewatering before it is discharged into an adjacent 20- by 8-ft.deep surge tank, where it has its first contact "with cyanide. To ensure good mixing the pulp is circulated by means of a 3-in. Wilfley pump, which runs continuously. From the surge tank the pulp (at about 35 pel' cent solids) is pumped to three 10-ft.-diameter by lO-ft.-deep Wallace agitators connected in series. Following a second filtration step with a water wash, the cake is repulped and agitated a second time in four 14-ft.-diameter by 14-ft.-deep Denver agitators, also connected in series. After a third filtration step using a barren-solution wash, the cake is passed to a 50-ft. final-washing Dorr thickener, overflowing pregnant solution while the underflow is sent to waste. To prevent solution fouling, approximately 40 tons of barren solution is wasted pel' day. About 700 tons of solution is clarified and precipitated in a Merrill-Crewe circuit per day (for details of precipitation and cleanup, see Chap; XI). OXIDIZED QUARTZ REEF

Marlu Gold Mining Areas, Ltd. (Type IIa). This plant, with installed capacity for 2000 tons per day, is treating 1000 to 1200 tons per day (1946) of outcrop material assaying 2.5 to 3.0 dwt. gold per ton. The are is the oxidized residuum of a network of small veins of quartz dispersed through an original shattered phyllite country rock. Most of the values are in the quartz, but some are in clayey associated decomposition products of the original phyllite. The lateritic overburden may also carry values which justify its inclusion in deliveries to the plant. The argillaceous components of the feed to the plant have poor settling and filtering characteristics. Following coarse crushing in a 48- by 60-in. Traylor crusher, the are is fed from twelve 250-ton storage bins through 12 washers similar to diamond gravel washers, fitted with lifters and l)/z- by YJ:6-in. trammel screens on the discharge ends. They revolve at 25 r.p.m., and solution sprays are installed in the discharge trammels to assist washing of the are. The oversize passes to two 7-ft. Newhouse crushers, which make a %-in. product that passes directly to four 6- by lO-ft. Head Wrightson ball mills closed-circuited with 7- by 30- by 14-ft.-diameter Dorr bowl classifiers. The classifiers, which also receive the washer trommel undersizs, overflow a pulp of 4.25 to 1 dilution carrying 80 to 85 per cent minus 200-mesh material. All washing and grinding are done in cyanide solution. The plant is of an all-sliming type and contains no traps or blankets. The pulp is next thickened in eight 75-ft. Dorr traction thickeners.

393


Mine ore

MilI solution storage

u-r.

1/

48 x 60 TrayJor jaw crusher

·f

3-000- ton ore bin

,-,

t

I " 7.

/I

12-4 x 16 Trommel woshers I Yz x '16 screens

r

t

(Fines)

(Coorse rock)

,

t f

2-7 Nell/house crushers

~ 4 -6',10'

-

neaa

Wnilhlson boll

mil"]

L--------==:::::j:4-7'x30'xI4' Dorr'bowl classifiers

t

.

,

(Overflow)

t

(Sands)

r-------- 8-75' Dorr traction thickeners f f

Pregnant solution

...---2

9- 30'x 10' Dorr agjtators • -SurQe tank ~ Banks of 8 vats 10'sqpare Butters filters

Barren solution

." t

,.-.

Sur9.e

,

agjtator~-------t

....------14 - 14 x 18 Oliver drum filters ...- - -....

+, ,

II

•

I

Cake to 15x 5 -6 surgfLg9.tfator To ml'll solution storage

,

r

t , -

30 x 10 Pull2 surQe tank 12-/4' x 18' Oliver filters

f

Cake rel2ull2ed

~Towaste 4-26-Leaf clarifiers

f f

Mem'll crowe I2reci(2italion ...... Barren solution storage Gold l2!ecig/fate FIG. 89. Flow sheet of the Marlu Gold Mining Areas .mill, Gold Coast Colony.

394


Lime (3.5 lb. per ton) is used to increase the settling rate. The use of starch flocculants, which were tried out, was discontinued owing to resulting underflows that were difficult to control. The thickener underflows pass to nine 30- by 20-ft. Dorr agitators, which operate at 40 per cent solids and 20 lb. per sq. in. air pressure. Owing to the high losses of dissolved gold in residues formerly experienced, a Butters filter plant has been installed to supplement the fourteen 14- by 18-ft. Oliver filters. Pulp from the agitators is first filtered in the Butters plant, the cake from which is repulped with barren solution to a dilution of 1 to 1 prior to delivery to the Oliver plant for final filtration and washing. Butters and Oliver filtrates pass to the circulating-solution storage. Thickener-solution overflows pass via clarifiers to the Merrill precipitation plant. Reagent consumptions are shown in Table 79. TABLE

79.

Aero brand cyanide Lime Zinc Lead nitrate

REAGENT CONSUMPTION AT MARLU

,

1.15 to 1.20 lb. per ton 3.5 lb. per ton are 0.07 lb. per ton are 0.01 lb. per ton are

BANKET ORES

Taquah and Abosso Mines, Ltd. (Type IIss). This plant is situated at Abosso in the Gold Coast Colony, on the banket lode. The ore is a dense, blue-gray, auriferous quartz conglomerate comprised of opaque white water-worn pebbles and disseminated grains of black hematite, the whole being compacted by secondary silicification to a homogeneous mass in appearance closely resembling the Witwatersrand banket in South Africa. The Gold Coast bankets contain no sulphides and in general are tougher than the South African. The gold is present in an extremely fine crystalline form, visible gold being. rare. The principal equipment units are indicated in the flow sheet (Fig. 90). Grinding is in cyanide solution. Twenty-five thousand tons is being milled per month (1948), and extensions to increase capacity to 35,000 tons per month are in hand. The average feed contains 5.0 dwt. per ton of gold, of which 2.3 dwt. per ton, representing 46.0 per cent, is recovered by amalgamation. Approximately 50 per cent of the total ore is cyanided as sand, assaying 2.75 dwt. per ton, and 50 per cent as slime (92 per cent minus 200 mesh), assaying 2.65 dwt, per ton. The total residue produced averages 0.26 dwt. per ton for an over-all gold recovery of 94.8 per cent. The consumption of forged-steel balls in milling is 3.14 lb. per ton. Reagent consumptions per tone ore are NaCN, 0.26 lb.; CaO, 2.5 lb.; zinc dust, 0.055 lb.; lead salts, 0.02 lb.

395


Run of mine are

M~ Cullevtusher (4~" seiJ f Magnet Tromp iron GriJ:- (t~'1

13"

Gxrex screen

-%" x fa" openings)

" -' +. (3.Vo'j -r::=..,o . unaersue versue

-

!

sxmotfJ(%~

4'

soml21i!Jf~P..lant_Head sample

S_t_o~-..oge (2,fOO tons)

.....

Prim ry batl mills 6'xlOj 2-Dorr dup..lex rake classifiers (6')

~

t

Underflow

ovejllow

2-Amalgaf plates

~-,---%-4--A...Jmalgam fJlates-----.......

Amalgam to retort. ~

T

Dorr bowlclassifiers (t4j

t

Ovedlaw

I

. t se~ndori batlmill(6'xI0; Underftow

Dorr bawl classifiers(IB;

f-

f

Sand

~

. -

.

Slime 2-ThickenJrs (50'x 10;

f

lo-Sutter 5 Mein distributors

Underflow 2-Pothucas (12'x 50')

, . t , ')

f (. , "

tv-LeocIi vols '35 xBf 'n

J

Overflow

2-Dorr agjlators(22x20 -2-Surge tanks125 X 10J

t

2~dliver

I.

" .

filers (14'x16')

SolLtian Merrill crowlP..!'eci,Qitolion

t iSmelt

Calcination

Sand residuehosed to dam

Slime residueto dam

FIG. 90. Flow sheet of the Taquah and Abosso mill, Gold Coast Colony.

396


BELGIAN CONGO

The gold and silver production of the Belgian Congo is partly derived from alluvial and eluvial workings in the east and northeast areas of the Congo Basin and partly as a by-product of the copper output. The principal operators in the Aruwini Ituri district in the northeast Congo are the Kilo Moto, the Miniere des Grand Lac, and Miniere de La Tele, while the latter companies also have holdings in the Kivu district to the south. The bullion is sent overseas for refining. Gold and black sands from the placers are shipped to central laboratories maintained by each company, where the former is melted and the latter amalgamated. The value of the gravels worked runs from an average of about 1 gram to as little as 0.25 gram per metric ton, which is considered to be about the lower economic limit. There are a few potential lode mining areas and two developed mines in the Kivu area which are milling a small tonnage of ore. Gold production fell during the war years from a peak of 561,030 oz. in 1941 to 381,206 oz. in 1945. Silver production reached its peak in 1942, with nearly 4,000,000 oz., falling to 2,500,000 oz. in 1945. TANGANYIKA TERRITORY

The Geita Gold Mining Co., Ltd., operates a 500-ton all-sliming cyanide plant in the Mwanza district. Plans are under way for increasing the plant capacity to 1000 tons per day. For the year ending June 30, 1946, 77,672 long tons were treated, yielding 15,525 oz. gold at a cost of 32.8s pel' ton. The New Saza Mines, Ltd., operates a 350-ton, all-sliming plant in the Lupa gold field and milled 77,232 tons during the year ending June 30, 1946, with the production of 15,679 oz. gold and 15,973 oz. silver, at a cost of 38.5 (E.A.) shillings pel' ton. KENYA

Small-scale mining operations are being carried out in various areas in this colony, of which those of the Kenya Consolidated Goldfields, Ltd., at Kitere in the Kavirondo district, yielding about 2000 oz. gold' in 19451946, and the 50-ton plant of the Rhamba Mines, Ltd., producing 3200 oz. gold, are typical.

SECTION 4.

AUSTRALIA, PACIFIC AREA, AND ASIA AUSTRALIA

This country is the fifth largest of the world's gold producers with a 1946 output of 824,000 oz. Its silver production for the same year (9 million ounces) was, however, about one-half the United States production. While the decline in gold output, which was in progress during the war,

397

TREATMENT OF GOLD ORES TABLE'

80.

PRINCIPAL PRODUCERS IN WESTERN AUSTRALIA*

Company

Big Bell Mines, Ltd.. ............. Blue Spec Leases ....... ','" ...... Boulder Perseverance ............. Broken Hill Pty., Ltd.......... " . Central Norseman Gold ........... Comet Gold Mines, Ltd ........... E dna May Amalg. G.M........... Emu Gold Mines, Ltd............. Evanston Gold N.L............... Gold Mines of Kalgoorlie ......... Golden Horse Shoe ............... Great Boulder Pty................ Hill 50 Gold Mine N.L ............ Kalgoorlie Enterprise ............. Lake View and Star, Ltd.......... Mountain View G.M .............. North Kalgurli (1912), Ltd........ Ora Banda Amalg................. Paringa Mining and Expl. ........ Phoenix a.M., Ltd................ Sons of Gwalia, Ltd ............... South Kalgurli Cons., Ltd......... Triton Gold Mine N.L ... , ........ Wiluna Gold Mines, Ltd .......... j

Where situated

Big Bell Nullagine Fimiston Kalgoorlie Norseman Marble Bar Westonia Agnew Evanston Kalgoorlie Boulder Fimiston Mt. Magnet Fimiston Fimiston Day Dawn Fimiston Ora Banda Fimiston Coolgardie Gwalia Fimiston Reedy Wiluna

Long tons treated, (2240 lb.)

357,623 4,029 137,456 44,307 107,750 oJ 2,768 17,498 26,720 11,585 158,337 .......

367,293 50,659 57,277 518,431 1,922 151,710 12,897 97,702 28,085 81,510 79,173 55,961 3,854

Gold Average No. produced.j of men fine oz. employed

.

41,048 4,645 33,498 13,893 34,411 3,744 6,774 6,247 5,019 39,138 10,648 94,151 13,673 17,807 148,766 12,795 44,609 3,704 21,429 6,785 24,986 19,503 14,382 11,101

384 59 354 133 362 24 61 71 63 327 46 821 104 118 834 16 278 67 200 80 235 177 260 122

* Excerpt from the annual report of the Executive Council of the Chamber of Mines of Western Australia, Inc., for the year 1947. t Ore produced by Boulder Perseverance, Ltd., Kalgoorlie Enterprise Mines, Ltd., and portion of that produced by North Kalgurli (1912), Ltd., were treated by Kalgurli Ore Treatment Co., Ltd. Ore produced by South Kalgurli Consolidated, Ltd., and portion of North Kalgurli (1912), Ltd., are were treated by Croesus Pty. Treatment Co., Ltd. A small proportion of the gold produced by Lake View and Star, Ltd., was recovered from tailings retreatment. Of the gold produced by Wiluna Gold Mines, Ltd., 825 oz. was obtained from the tonnage treated, the remainder being production from cleanup of plant and retreatment of tailings. Gold production by the Norseman gold mine ceased at the end of January, 1947. This company is now producing pyrite. appeared to have been halted in 1945, high costs and labor shortages had restricted a revival of the industry. Many important developments in the metallurgy of gold and silver ores

398


have originated in Australia, which is noted for the skill of its metallurgists in solving difficult ore-treatment problems. WESTERN AUSTRALIA

Western Australia is by far the most important gold-producing state in the Commonwealth. The area in and around the famous Kalgoorlie field has within a radius of 12 miles produced more than half of Western Australia's output to date of 1,556 long tons of gold. Since its discovery in 1893, this field has yielded about 27 million ounces from the milling of 50 million tons of ore. The bromocyanide process, now abandoned, was widely used in this district for the treatment of the refractory sulphotelluride ores. Many of the important producers, such as Hannan's Star, Oroya-Brown Hill, and Golden Horseshoe Companies, have since closed down, but the extensive retreatment of tailings from earlier operations has helped to keep the district active. The 1946 production of around 600,000 oz. was about one-half the prewar figure, due to such handicaps as labor, equipment, and supply shortages. The ore of the "Golden Mile" near Kalgoorlie .oecurs in quartz-dolerite greenstones and calcite schists with occasional albite or hornblende porphyry dikes. There is little or no demarcation between country rock and ore, only careful sampling revealing the lode limits. The ore is fairly hard and carries an average of 5 per cent pyrite and varying percentages of tellurides. STRAIGHT CYANIDATION

Big Bell (Type IIa).24 This is the largest low-grade milling operation in Australia and the only large plant in which countercurrent decantation is used. The present capacity is about 1200 short tons per day of ore running less than 2.8 dwt. per ton. Coarse crushing is carriedout in two 36- by 24-in. Ruwolt jaw crushers breaking to 3 mesh. Secondary crushing is done in two 4-ft.-diameter Symons cone crushers, and approximately a ~-in. product is delivered to the fine ore bins. The grinding section consists of 8- by 6-ft. Ruwolt primary ball mills closed-circuited with 26-ft. 8-in. by 8-ft..Dorr classifiers. One-half the rake sands is reground in secondary ball mills of the same size, and the pulp classified in bowl classifiers to overflow at about 1.5per cent plus 65 mesh (52.7 per cent minus 200 mesh). All ball mills revolve at 22 r.p.m. and carry a 6:1 circulating load. The ball charges are 17 tons of 4- and 2in. steel balls. Cyanide solution is added at the primary mills, and lime at the secondary 24

Excerpts from article in C.E. and M. Rev., Mar. 10, 1941, pp. 166-170.


399

mills. Automatic density controllers are installed on the bowl classifiers, the overflow of which is thickened in a 100- by 14-ft.-deep DO~T Torq thickener, which overflows clear pregnant solution for precipitation in a Merrill-Crowe plant. . Agitation is carried out at 45 per cent solids in two parallel banks cit five 30-it.-diameter by 18-it.-high agitators, and aeration is maintained by 28 lb. per sq. in. low-pressure air, supplemented by high-pressure air from the mine supply as needed. The countercurrent decantation washing is carried out in five thickeners, with barren solution added at thickener 4 and wash water at thickener 5. The final underflow is discharged at 52 per cent" solids. The consumption of water in the mill is 320,000 gal. per 24 hr. The area of the agitators, tanks, and thickeners is 57,000 sq. ft., and the loss of water at an average evaporation of 0.35 in. per day.is 11,000 gal. per day. The consumption of stores in pounds per ton of ore treated is 4-in. balls, 0.952; 2-in. balls, 2.179; cyanide, 1.010; lime, 4.224; zinc dust, 0.181; lead acetate, 0.018; litharge, 0.75. CONCENTRATION AND CYANIDATION

C. Blackett, well-known consulting metallurgist, writes as follows concerning present practice in western Australia: The principal plants using straight flotation, calcining of the concentrates with subsequent cyanidation are Lake View and Star and Croesus Pty. Treatment Co., Ltd. The latter plant is the newest and most up to date in this category and is the only plant that has installed a Lodge Cottrel dust precipitator. The Kalgurli Ore Treatment Co., Ltd., is the best example of precyanidation, activating the cyanide residue with copper sulphate and soda ash before flotation. The Gold Mines of Kalgoorlie, Ltd., and Great Boulder Pty., Ltd., follow Lake Shore practice, where the cyanide residue is conditioned with S02 gas from the roasting furnaces. South Kalgurli Consolidated, Ltd., abandoned the all-dry-crushing and roasting plant some years ago, and the are from this mine is now being treated at Croesus Pty. Treatment Co., Ltd., together with a portion of North Kalgurli (1912), Ltd.

Lake View and Star, Ltd. (Type IVr).25 The present mill, known as the Chaffers ore-treatment plant, was adapted from a dry-crushing and all-roasting process plant after the present company had successfully operated a pilot section utilizing flotation for the first time in Kalgoorlie for the treatment of a pyrite gold ore. Its capacity was increased by stages as the ore development warranted, each new grinding unit taking the place of a dry-crushing mill until the latter were all discarded. 26

From Bul. A.I.M. and M., May, '1948, compiled for the first ordinary meeting.

· 400


Its maximum capacity at the present time is 60,000 short tons per 28day period, but this figure has not been attained since the war. The present throughput is 45,000 short tons per 28 days, labor shortage underground being the reason for the reduced tonnage treated. The process' used consists of crushing and fine grinding followed by flotation of the auriferous pyrite and tellurides. The flotation concentrate so produced is then roasted and cyanided. Subsequent cyanidation of the flotation tailings, which are at present discarded, is now under consideration and will be shortly given a test run utilizing the existing Chaffers retreatment plant which is described later. Crushing is done on two shifts from Monday to Friday inclusive. All subsequent operation are continuous for 24 hr. each day. Crushing Plant. Primary crushing is done at two stations by 3-ft. Traylor gyratory crushers. One station is situated at the Chaffers main shaft and is fed from the shaft bin of 100-ton capacity. The second station receives all ore received at the plant by surface transport from the Ivanhoe, Lake View, and Associated shafts. This ore is transported to the crusher station by Diesel-locomotive hauling rakes of ten 4-ton trucks which dump into two 100-ton bins situated on either side of the Traylor crusher. The primary crushers reduce the run of the mine ore to 1~i-in. size, the crusher discharge being carried by inclined conveyors to the central secondary crushing plant. Electromagnets recover tramp iron prior to screening, which removes the minus .%-in. ore from the feed to the two standard 4-ft. Symons cone crushers. These perform the secondary crushing and use fine-setting bowl liners. The secondary crusher discharge joins the primary screen undersize and is elevated by conveyors to the four finishing screens, using %-in.-aperture wire-mesh screening. These screens are situated over a 250-live-ton storage bin which receives the oversize from the finishing screens. This oversize material is then fed by two 36-in. ribbon feeders to the tertiary crushing units, which are 4-ft. Symons short-head crushers. The discharge from the short heads joins the secondary-crusher discharge and is returned to the finishing screens to produce a closed-circuit crushing operation. The minus %-in. undersize is carried by an overhead conveyor to either the mill running bin or to the storage bin system. The crushing-plant capacity is 1400 tons per 8-hr. shift. Plant bin storage consists of One 6600-ton circular steel storage bin.

TREATMENT OF GOLD

ORE~

401

One 3000-ton rectangular wooden storage bin. One 1100-ton wooden mill running bin. AJI ore on its way to the running bin is sampled automatically and passes over a Merrick weightometer. . Grinding Section. Grinding is done in three stages prior to flotation followed by a tailings regrind. By means of shaking chute feeders, 6- by 5-ft. primary ball mills (six off) take the ore direct from the bin. The mills are of the trunnion-discharge type using manganese-steel shiplap liners and fed with mixed 4- and 3-in.-diameter forged-steel balls. The average ball load is 7 tons. oJ The mills are open-circuited, and the discharge is passed over corduroy cloths for the recovery of free gold. These cloths are changed every 12 hr. and washed, the product going to an amalgam barrel. The recovered amalgam is treated in the gold room. After passing over the primary strakes, the ball-mill product is classified in 25- by 6-ft. DOlT duplex classifiers. The classifier sands are ground in the secondary mills, the work being performed by three only 5-ft. 6-in. by l1-ft. and two only 5-ft. 6-in. by 22-ft. tube mills. Tube-mill discharge is returned in closed circuit to the primary strakes and then back to the classifiers. The strake area is 11 tons per sq. ft. of circulating feed. The 5-ft. 6-in.-diameter mills are of the grate-discharge type, using 2-in. cast-iron balls and lined with modified EI Oro liners. The ball load is l1-ft. mills, 11 tons; 22-ft. mills, 13 tons. Classifier overflows combine and are pumped over scavenging strakes to two 14-ft. bowl classifiers. The secondary strake area is 6.6 tons per sq. ft. The sands from the 14-ft. bowl classifiers are ground in open circuit in a 5-ft. 6-in. by l1-ft. tertiary mill, the discharge of this mill joining the bowl-classifier overflow in the flotation feed surge tank. The tails regrind section is described under "Flotation." Flotation. Five "rougher" machines handle the feed from the flotation surge tank, each machine having 10 size 24 Denver Sub-A cells. The product from the first cell is pumped directly to the concentrates thickeners, the overflows from the remaining nine cells being cleaned in a six-cell South Mine flotation machine or a Ill-cell mineral separation Sub-A machine. All concentrates from the "cleaner" tails after thickening are returned to the secondary mills. "Rougher" tails are reclassified in a 20-ft. QSFB classifier, the overflow being the finished tailing, and the sands are reground in a 5-ft. 6-in. by l1-ft. mill. The mill discharge is returned to the flotation surge tank. Finished tailings are thickened to 59 per cent solids and pumped to the dam. Clean concentrates after thickening are dewatered on a 8- by 5-ft.

402


3-in.-diameter Oliver filter.

The ratio of concentration by flotation

IS

127~: 1.

Average sulphur content of concentrates is 36 per cent. Roasting. The dewatered concentrates are passed over a Merrick weightometer into a 8-ton bin from which they are fed by a ribbon feeder on to the main roaster feed belt (see Ohap. X). The roaster discharge is water washed to remove soluble iron sulphates. In the water-wash circuit the calcines are straked to recover the free gold liberated in the roasting by the decomposition of the tellurides contained in the concentrates and given ,a closed-circuit grind in a 4- by 4-ft. mill and 20-ft. by 6-in..classifier. The strake area is 1 ton per sq. ft. of calcines. The strake cloths are treated similarly to those from the mill strakes but in a separate amalgam barrel. Cyanidation of Calcines. The water-washed calcines after dewatering on an Oliver filter are repulped with cyanide solution and pumped to the agitating section, which consists of five Dorr agitators coupled together in series. Agitation time given is 80 hr. in a 34 per cent solids pulp. The cyanide strength is maintained at 0.08 per cent KON, and the lime strength at 0.02 per cent OaO. The pulp leaving the agitators is given two stages of decantation followed by vacuum filtration and washing on an Oliver filter. The filter cake is then discarded to the dam. Clarification and Precipitation. All pregnant solutions leaving the decantation and final filtering stage are clarified in a converted Cassell filter using double thickness of calico sheeting as the filtering media. The clarified solution is then de-aerated, and the gold recovered in a 1400ton vacuum-leaf-type Merrill-Orowe precipitator using zinc dust. The barren solution is measured through a Kent integrating and recording meter and returned to the cyanide section for further use. Smelting of Gold Slime. The gold-zinc slimes are recovered from the Merrill precipitator three times per 28-day period and after pressing dry in a small Dehne press are roasted, mixed with the requisite fluxes, and smelted in a Wabi bullion furnace. The fluxes used include pyrites from the flotation plant, and refining of the bullion is done in these furnaces by the production of a copper-lead matte. The bullion produced is remelted and sent to the Perth Royal Mint. The slag is amalgamated and cyanided on the mine, and the matte re-treated by a highly reducing charge to form a base bullion, which collects the gold and silver values. This base bullion is then cupeled and dispatched to the Royal Mint.


403

Chaffers Re-treatment Plant. This plant was designed and erected to treat a dump consisting of roasted-ore residues and the early flotation tailings, situated % mile to the south of the Chaffers ore-treatment plant. The dump was estimated to contain 5 million tons of treatable material that averaged 1 dwt. per ton. The plant capacity is 55,000 short tons per 28 days. . The residues are sluiced to the plant by hydraulic monitors serviced by a two-stage 5- by 7-in. centrifugal pump. This pump is capable of sluicing to the agitators the equivalent of 100 tons per hr. of dry slime in a pulp of 50 per cent solids. Glass nozzles 1% in. in diameter are used, "having an average life of 5000 tons. The high specific gravity for agitation is obtained by returning slime through the monitor until the desired gravity is reached and then passing a portion to the agitator filling pump while the rest, with make-up solution, is returned to the sluicing pump. TABLE

81.

TYPICAL METALLURGICAL RETURN: LAKE

Vmw

AND STAR

Percentage of gold recovered: By mill strakes :. . . . . . . . . . . . . 17.79 pel' cent By calcine strakes. . . . . . . . . . . . . . . . . . . .. 19.71 By cyanide plant...... . . . . . . . . . . . . . . . .. 54.46 Decrease in gold in process. . . . . . . . . . . . . .. -1. 49 Total. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. 90.47 per cent

Dissolving is done in Devereaux agitators part in series and part batch, The batch agitators follow the series ones and are used for calculating tonnage hancUed. A portion of the dissolving is done during sluicing operations. Time of contact in agitators is 4 hr.; cyanide strength, 0.03 per cent KCN; lime, 0.002 per cent CaO. Four only 16- by 14-ft. Oliver vacuum filters hancUe the agitated slimes in the filtering and washing section, the filter cake being repulped with waste water from the Chaffers plant calcine wash section and sent to the dam, while the filtrate is clarified prior to precipitation. The filtering area is 2800 sq. ft.; the filtering rate, 1360 lb. per sq. ft. per day of dry slime. Clarification of the solutions is carried out in four leaf-type converted Cassell filters which handle up to 70,000 tons of solution per 28 days. Precipitation of the gold using zinc dust is done in two 1400-ton MerrillCrowe precipitators of the vacuum-leaf type. Barren solution is metered by two Kent integrating and recording meters and returned for use in the 'sluicing and filtering sections. The gold-zinc slimes are cleaned up fortnightly and after pressing are taken to the Chaffers plant gold room for smelting. No preliminary roast

404


is given, and the slimes are mixed with an oxidizing flux and smelted in the Wabi furnaces to produce bullion and slag. The Croesus Proprietary Treatment Co., Ltd. (Type IVb).26 was formed on Mar. 12, 1945, for the purpose of treating ores from South Kalgurli Consolidated, Ltd., and North Kalgurli (1912), Ltd., on an equaltonnage basis. Until March, 1948, the capacity of the plant was approximately 7000 long tons pel' 28 days, but quite recently the addition of a new grinding, straking, and flotation section coupled with an additional roaster and sunken agitator has increased the throughput considerably, and during the period ended May 25, 1948, a total tonnage of 11,978 long tons was realized. In general the ores are quartz dolerite greenstone and calc schist, with occasional albite or hornblende porphyry dikes. Broadly speaking the treatment consists of a closed-circuit crushing with a 30- by 18-in. Ruwolt jaw crusher and a 4-ft. standard Symons cpne crusher in conjunction with a 48- by 102-in. heavy-type suspended Gyrex carrying cloth of 17~2-in. mesh. A series of coarse-ore storage bins is necessary to keep the parcels of the two companies separate until sampled and weighed. Fine grinding is effected by an 8- by 5-ft. Fraser and Chalmers mill and a 57~- by 11-ft. Ruwolt unit. Classification is done with Dorr bowl and rake machines. In the grinding section 15 strake tables 7 by 2% ft. are available. The concentrates produced from two eight-cell banks of the Denver Sub-A type in parallel are roasted in two Edwards duplex 54-rabble roasters, and the resulting calcines are cyanided in five sunken agitators, using the batch system. A modern Merrill-Crewe unit handles the solutions produced by an 8- by 12-ft. Oliver filter. The gold-zinc slime accumulated by a cornerfeed-type Merrill press is cleaned up twice per period. Features of this plant are the Lodge-Cottrell precipitator operating on the roaster flue gases and a splendid sample mill which is a definite asset when ores from two companies are accepted. Flotation samples are taken automatically, and the time of cut mechanically recorded. Safety devices and alarms on mill solutions and ore feeders have been installed throughout the plant, and a general modern aspect is apparent, from crude ore to the finished product. The Kalgurli Ore Treatment Company (Type IIIr).z7 Each of three contributing companies delivers its coarse ore into a corase ore bin, from which it is drawn to pass over a stationary grizzly set at IH~in. opening and the oversize fed into a 3-ft. Traylor crusher also set at IH-in. Private communication from C. Blackett. Paper read at the May, 1948, meeting of the Australian I.M.M. in Western Australia. 26

27

405


Mine ore bins

-r-

t

2-ln Grlzz/x...

r

. . 30 "x la " Ruwoll jaw crusher 102"x 4a" Gy.!ex vlbratlna screen (/~2" ctotn)

4 fl. slandard cone crusher I

Chain and Buckelsaml2/er - - ,

l

-.,

Fine ore bins "

I

Sample mill ~

_w,_eI9imeler

I

>(

-,-

8 x 6 Fraser 8 Chalmers ball mill . 5 Yz"x II "Ruwoll ball mill

Sample

10 a~say

office

!bleS(2~9"X 7')

16-Slrake t---l~fS§..J..t Concenlrate amalgamaled

a' Dorr Duplex classifier 16'Dorr quadrul2.lex bowlclassla' dlam.

f

-

f

(Overlow)

(Rake)

6'x6' Denverconditioner

' - - - -.....

t

2-Banks No.24 Denverflotatloner cells

f

f

tconc.)

tlOlIs) 70' Dort torq. thickener dam.

ur: to tailings

20' DorrftorC/: thickener


~

6- Americandisc filters

~

2-64-rabble Edwards roasters

l 4-20'x a' Goldfield /,Yl2.e agitators (2 air-lifts) ,

, t

a x 12 Oliver filler

Filtrate

l

.-

20'x a' Goldfield tY/2.e ag./lator ~--+------t -a'x12' Oliver filter Filtrate

~ Cake to residue dam

Clarification andprecipitation bullion

FIG. 91. Flow sheet of the Croesus Proprietary mill, Western Australia.

406

CYANIDA~ION

AND CONCENTRATION OF ORES

opening. From here the ore passes into individual crushed-ore bins, a separate bin for each company. When convenient, ore is taken from the various bins, screened over four 8- by 4-ft. gyrex screens, the undersize passing to the weightometer while the oversize is carried by conveyors past two electomagnets to remove tramp iron and into a 5-ft. 6-in. Symons short-head cone crusher set with a minimum opening of % in. The crushed ore then passes over the same gyrex screens as the fresh feed, the minus % in. going on to the weightometer, while any oversize returns to the crusher for further reduction. The crushed ore is weighed by passing over a Blake-Dennison automatic weightometer, approximately 400 tons of any individual company's ore being taken for a parcel. The weighed ore is then sampled, 4 per cent being removed by means of a vezin sampler reduced to minus 17 mesh in a No.8 Krupp ball mill and reduced to a sample suitable for assay. After being sampled, the ore passes on belt conveyors to the mill feed bins, no further care being taken to keep the companies' ore separate as the tare and grade are now known. It may be mentioned that, as usual, buyer, seller, and umpire samples are kept, the latter being assayed only if agreement is not reached between the buyer and seller. The reject from the Krupp Ball Mills also passes into the mill circuit, being fed into the distributor at the head of the secondary corduroy tables. From the mill bins the crushed ore is fed continuously, at a rate dependent on the mill circuit, into the three primary ball mills where it is ground in a cyanide solution. The three ball mills consist of an 8 by 6 Hardinge mill, a 6 by 5 Ruwolt mill, and an 8-ft. fi-in. by 6-in. Marcy grate discharge mill using equal weights of 3- and 4-in. steel balls as grinding mediums. After passing through the primary ball mills, the coarse slurry is pumped to a distributor and over twelve 7- by 3-ft. corduroy strake tables and then divided into a 16- by 2-ft. quadruplex classifier, and an 8-ft. by 23-ft. 4-in. multizone classifier, and a portion into the secondary circuit by way of the secondary classifiers. . The oversize from the classifiers is returned to the primary mills in the case of the primary classifiers; and to the secondary mills in the case of the secondary classifiers. The undersized product overflows and reaches the secondary circuit by way of an 8-ft. by 31-ft. 8-in. by 16-ft. Dorr bowl classifier, where it is reclassified, the oversize going to a ll-ft. by 5-ft. 6-in. Ruwolt secondary mill while the undersize overflows the bowl periphery as the final ground product for thickening. The secondary grinding circuit contains two l l-ft. by 5-ft. G-in. Ruwolt mills, and these work in closed circuit with twelve 7- by 3-ft. corduroy strake tables and a 16- by 20.:.ft. quadruplex classifier, as well as the bowl classifier as was previously mentioned. The product from both primary and secondary strakes is washed

407

TREATMENT OF GOLD ORES Coarseore bins

t

t.f2" StatlonafLgrlzzif-

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t -.,

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t

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FIG. 92. Flow sheet of the Kalgurli Ore Treatment Company mill in Western Australia.

408


once per day into two 4- by 2-ft. amalgamating barrels and ground with mercury, the resulting amalgam being subsequently collected and treated in the gold room. After the grinding is completed, the bowl-classifier overflow pulp contains 20 per cent solids, which have a grading of 75 per cent minus 200 mesh. This pulp forms the feed to five 38- by 8-ft. Dorr thickeners, which thicken the pulp to 52 pel' cent solids, while the clear overflow is returned to the mill-room cyanide head tank. The thickened pulp is pumped by means of Dorrco pumps into agitator vats where it is agitated at 0.04 per cent N aCN and is passed along by air lifts between each vat through a series of six vats. From the last vat the pulp is pumped into a surge tank and distributed into three 12- by 8-ft. Oliver filters, one 12- by 8-ft. Paxman filter, and one 16- by 14-in. Oliver filter, the cake being washed with water to give a final soluble N aCN value in the cake of 0.002 to 0.003 pel' cent NaCN. The filtrate from the filters is pumped into a turbid solution tank and from here is pumped into five 30- by 4-ft. sand clarifiers and one 40- by 4-ft. sand clarifier to give a clear solution before precipitation. From the sand filters the clear solution gravitates to a pregnant solution tank and is pumped into a Merrill-Crewe vacuum tower to remove dissolved oxygen. Zinc dust is then fed into the deoxygenated solution to precipitate the gold and the whole pumped into two 52-in. Merrill presses. Here the gold slime is filtered out, while the barren solution passes from the presses into a tail solution tank from which it is pumped back to the mill-room head tank and the calcine section head tank. The gold slime is cleaned up three times per period and is treated in the gold room. Returning to the Oliver filter cake, after washing this is discharged into vortex mixers for repulping with solution from the flotation circuit. The mixers discharge into an agitator where the pulp is activated with 0.5 lb. of copper sulphate and 0.75 lb. of soda ash per ton of ore. From this agitator the pulp is pumped into a conditioning tank before flotation where 0.2 lb. per ton of sodium secondary butyl xanthate is added and the pulp gravitated into three banks, of 10 units each, of Fahrenwald flotation cells. As a frother, equal quantities of pine oil and cresylic acid are used. The temperature of the float is kept at 33 to 35°C. The flotation tailings at approximately 38 per cent so ids discharge into a 100-ft. traction thickener where the pulp is thickened to 60 per cent solids and pumped to a tailings dam. The overflow from the thickener is pumped through circulating pipes in the roasters for heating the solution and is used for repulping the cake from the Oliver filters.


409

The concentrate from the. flotation cells is thickened in a 25- by 8-ft. concentrate thickener and discharged into a concentrate agitator previous to filtering. The overflow from this thickener contains a quantity of slimed pyrite and tellurides which are extremely difficult to thicken, and this overflow is refloated through two unit Fahrenwald flotation cells. The concentrate joins the thickened concentrates in the agitator, while the tailings pass again into the flotation circuit. The thickened concentrates are filtered in an American filter having three sets of 10 leaves. The moisture content of the filtered cake is from 8 to 10 per cent moisture, and in this state it is discharged onto three belts which feed directly into three 56-rabble Edwarcts roasters. The concentrate is dried by the outgoing gases and catches alight at the 4 to 5 rabble and roasts through the various stages of pyrrhotite, magnetite, and hematite until completely roasted. Covering several rabbles from the center of the roasters, sets of five I-in. water pipes run transversely in the roaster gas both to cool the roast at that point and to heat up the' 100-ft. thickener overflow solution for raising the temperature of the flotation pulp. The roaster gases are cleaned in a Buell cyclone dust collector, the dust being returned in a screw conveyor to the roasters while the gases are discharged through a 200-ft. steel stack. The hot calcine from the roasters discharges into a push-conveyor set across the discharge of the three roasters and is taken by another push-conveyor to an elevator. The calcine cools in the push-conveyors and from the elevator is discharged into the first of 'a series of six 20- by 8-ft. agitator vats. The calcine is pulped with barren solution and agitated for 70 hr. The pulp is filtered three times through a 12- by 8-ft. Oliver filter, the cake being repulped and agitated between each filtration. Finally the cake washed with barren cyanide solution is repulped in a vortex mixer and pumped to a calcine residue dam. The filtrate from each filtering is pumped to the turbid tank mentioned previously and joins the filtrate from the raw-ore filtration. In the gold room the gold slime from the Merrill presses is roasted, smelted, and cast into bars of bullion for banking. The amalgam from the corduroy strakes is retorted and the resultant gold cast into bars, also for banking. Recovery by strakes . . . . . . . . . . . . . . . .. 9 per cent Recovery by precyanidation 42 Recovery by flotation 42.5 Total recovery. . . . . . . . . . . . . . . . . . . .. 93.5 per cent Treatment costs for 1947: 14s4d per long ton, or 12s9d per short ton. These costs include overhead, general charges, bullion charges, etc.

410


Gold Mines of Kalgoorlie, Ltd. (Type IIIr).28 This plant, which is situated on the eastern edge of the "Golden Mile," originally had a capacityof 9000 tons all-sulphide ore per month, but it was extended in October, 1939, to treat 13,000 to 14,000 tons per month of mixed sulphide and oxidized ore. ~M ethod of Treatment. The treatment consists of fine grinding and cyanide treatment followed by flotation of the cyanided residues. The flotation concentrate is roasted and then cyanided. The gold is recovered from the cyanide solutions. The Ore. The mill feed is now made up of Per Cent

Australia East open-cut ore Oroya south mine ore , .. Iron Duke are. . . . . . . . . . . . . . . . . . . . . . . . . . .. New North Boulder ore

36 25 22 17

The open-cut oxidized ore consists mainly of quartz dolerite greenstone with a fair proportion of clay and old stope filling, the latter being roasted and unroasted tailing. The open-cut sulphide ore is chiefly quartz dolerite. The Iron Duke ore is all sulphide and is of two types: calc schist and quartz dolerite. At times a small proportion of sediment is mixed with the calc schist, and this sediment is, in part, graphitic. Sulphide mineral, practically all pyrite, occurs to the extent of approximately 6 per cent in the Iron Duke ore and 2 to 3 per cent elsewhere. An analysis of a 6 months' bulk mill feed sample gave the following results: TABLE

82.

CHEMICAL ANALYSIS OF MILL FEED: GOLD MINES OF KALGOORLIE,

LTD.

Gold (Au) Silver (Ag) Copper (Cu) Sulphur (S) Lead (Pb) Zinc (Zn) Arsenic (As) Antimony (Sb)

5.6 dwt. per ton 2.4 dwt. per ton 0.03 per cent 2.24 per cent 0.005 per cent 0.21 per cent 0.013 per cent 0.005 per cent

Metallurgy. From an operating point of view, the main metallurgical problems are 1. The problem common to all Golden Mile ores, the lockup of a portion of the gold in an extremely fine state, in some of the pyrite, even after grinding the pyrite to all minus 200-mesh screen. 28 Paper prepared for the Australasian Institution of Mining and Metallurgy, first ordinary meeting, May, 1948.


411

2. The wide variations that occur in the mixture of ores and stope filling received for mill feed. . 3. The Iron Duke graphitic material, which is received at irregular intervals and varies in type and in quantity. . These problems resolve themselves as follows: Seventy-five per cent of the gold is recovered from the crude ore in a precyanide section by fine grinding and straking and by further agitation treatment in cyanide solution. The remaining 25 per cent is chiefly locked up in pyrite particles. A pyrite gold concentrate is recovered by flotation and subsequently roasted, the pyrite particles bJeing converted to porous grains of ferric oxide. Being porous, the greater portion of the encased gold is exposed and recovered by further cyanide treatment. The problem of variations in the mill feed mixture principally affects operation in the grinding section. At times, when the feed changes from practically all-oxidized to all-sulphide in the matter of a few minutes, millmen have to be very much on the alert to avoid overfilling" a mill 01' acquiring risky, high-circulating loads. Flotation is the section most affected by excessive quantities of oxidized ore, and operating control is most important. The flotation has to be watched closely to see that a stable froth is maintained and to avoid oxidized slime, floating with the concentrate. The "graphitic" material from the Iron Duke mine, which is treated from time to time, does not seriously affect recovery provided only a reasonably small amount is present in the ore. To this end, care is taken in ore selection underground and in sorting it from the primary conveyor belt in the crushing section. Flotation Section. The repulped pre cyanide residue, together with flotation middlings, is pumped by the gas tower feed pump to wooden towers 30 ft. high and 4 ft. 6 in. square. Only one tower is used at a time, the towers being changed once a week for routine cleaning. Pulp density is regulated to 52 pel' cent solids. Some of the S02 exit gas from the Edwards roasters is blown in at the bottom of the tower and meets the stream of pulp, cascading down, over baffles, inside the tower. Sulphur dioxide gas is absorbed by the pulp and subsequently mainly converted to sulphuric acid by extremely vigorous agitation and aeration in three specially adapted Devereux agitators. As a result, the pH of the solution drops from 8.0-10.0 to 4.0-4.5. Carbonate in the pulp is dissolved by the acid, causing the pH to rise slowly, and as the S02 content of the pulp diminishes to nil, the pH reaches 6.8 to 7.2. However, the pH is low sufficiently long for the acid to accomplish the cleaning of the mineral surfaces and change the mineral from a slimed condition to a clean, readily flotated condition. Without this acid treat-

412


ment, the concentrate is slimy and low grade and recovery of pyrite and gold is poor. During the 6 hour treatment in the Devereux agitators, the dissolved gold in the pulp, 0.2 dwt. per ton, is precipitated and subsequently recovered with the flotation concentrate. 83.

TABLE

FLOTATION SECTION ROUTINE ASSAYS

Gold, dwt. Dissolved per ton gold, dwt,

Sample

Precyanide residues. " .............................. Flotation feed ....................................... Concentrates ........................................ Middlings ........................................... Tailings (including calcine tails) ..................... Tailings (excluding calcine calculated at 3 dwt. per ton) . ............................................. TABLE

84.

Sulphide sulphur, per cent

1.4 1.8 24.8 4.0 0.39

0.20 .... ....

1. 92 2.48 43.0

0.01

0.17

0.30

0.01

FLOTATION REAGENT CONSUMP,TION

Copper sulphate. . . . . . . . . . . . . . . . . . . . . .. 0.35 lb. per ton milled Ethyl kanthate 0.08 Butyl xanthate 0.08 Cresylic acid :.0.04 Pine oil and eucalyptus. . . . . . . . . . . . . . .. 0.01 TABLE

85.

RECENT 6-MONTH BULK SAMPLE OF CONCENTRATE

Gold (Au) Silver (Ag). . . . . .. . . . . . . . . . . . . . . . . . . . .. Lead (Pb)............................. Zinc (Zn).............................. Iron (Fe) Sulphur (S) Copper (Cu).... . .. .. . . . .. . . . .. . . .. . Arsenic (As) : Antimony (Sb).

24.8 dwt. per ton 11.2 dwt. per ton 0.02 per cent 0.39 41.1 43.02 0.33 0.17 0.05

Following a further I-hr. conditioning with xanthate at 40 per cent solids, the pulp is floated in two 12-cell Denver flotation machines using copper sulphate, cresylic acid, and sodium butyl xanthate. Roasting and Calcine Treatment. Control of the draft on each roaster is mainly by operation of dampers and offtakes partway down the furnace, and dust collected by the multicyclones is fed back to the roasters by screw conveyors. Gas for the 802 process is taken off at the main offtake by means of a small fan. On account of the high sulphur content of the concentrate, no extraneous


413

fuel is required, and once the charge has been ignited, using oil burners for the purpose, it continues to roast indefinitely as long as the proper feed rate is maintained. The calcine is pumped once per shift to two Devereux-typo agitators. In the first it is pre-aerated in a lime solution (0.01 pel' cent CaO), and-in the second it is cyanided for 48 hr. at 45 pel' cent solids using 0.06 per cent cyanide (KCN) and 0.01 pel' cent lime (CaO) solution strength with the addition of 0.02 lb. per ton of lead nitrate. The cyanided residue is returned to the precyanide mill circuit to receive further grinding, straking, and cyanide treatment, and then an acid treatment in the S02 process, followed by flotation treatment. It is the experience of the operators that when this is done the calcine residue is substantially reduced in value, particularly when it is higher than normal, e.g., when it assays 4 to 8 dwt. pel' ton. The total cyanide consumption in the treatment of ore and calcine is 1.10 lb. NaCN per ton of ore milled; the total lime consumption 5 lb. per ton. The 80 2 Process. Tests have shown that from 20 to 30 lb. sulphuric acid pel' ton of ore is required to condition the plant pulp satisfactorily so that reasonable flotation can be obtained. The S02 provides an extremely cheap source of this acid and makes the treatment of these oxidized ores economical. It has been established that, if pure pyrite mineral is pulverized without contamination and agitated with a pure solution of gold in cyanide solution, the whole of the gold is precipitated in a reasonable time and can be filtered or floated off with the mineral, leaving a barren solution. In the process described, after gassing with S02, the mineral in the pulp is cleaned by the acid in the Devereux agitators and is thus enabled to pick up for subsequent flotation the dissolved gold which has escaped in the precyanide-section filter residue. Control of the gassing is most important and is obtained by regulating a butterfly damper in the gas flue. The density of the pulp delivered to the gas tower is kept fairly close to the set figure, and a pyrometer, located in the gas flue, indicates by changes of temperature when gas changes, outside the control of the' flotation operator, are taking place. Samples of 50 cc of solution are filtered hourly from pulp samples taken at the gas towel' discharge; the Devereux agitator discharges and the flotation feed pulp are titrated with O.lON iodine solution and provide a very good means of control. The main object is to use as much S02 gas as possible provided the subsequent Devereux agitation reduces the S02 content to at least 0.4 cc iodine by the time the pulp reaches the flotation machine.

414


The advantages of the process are summarized as follows: 1. It permits of the economical treatment of large percentages of oxidized ore. 2. It reduces dissolved losses to 0.01 to 0.02 dwt. per ton. 3. Some extra gold is recovered from the calcine residue, particularly when it is of high value. 4. It eliminates the costly separate filtration of calcine, with its attendant high dissolved gold losses. 5. It reduces flotation reagent consumption by at least 50 per cent. 6. Not having to worry about increased dissolved losses in the precyanide residue, it was possible to alter the original thickening circuit from two stages of washing thickening to one and thus use the surplus thickeners to increase the capacity in primary thickening and one stage of washing thickening. The thicker pulp obtained increased filter efficiency, so that both thickening and filter sections were capable of handling greater tonnages from the grinding section. VICTORIA

Important dredging operations have been carried out in this Australian state for many years. The Lodden River operation of the Victoria Gold Dredging Co. has handled a total of 18 million cubic yards of gravel since 1938 with an average yield of 2.97 grams per cu. yd., or a total of 1~1,779 oz. gold. Recent figures for the two important producers of this area are shown in Table 86. In addition to dredging, it is reported that there are 40 small cyanide plants in operation in the state, the largest of which is the Rutherglen Gold Dumps. The outstanding gold mines of the state are the A1 Consolidated at Gaffney's Creek, which produced 18,601 oz. in 1946-1947, and the Morning Star mines. The gold yield for Victoria in 1946 was about 87,000 oz. TABLE

86.

GOLD PRODUCTION IN VICTORIA,

1947

Victoria gold dredging 2,111,457 cu. yd., 10,737 oz. Au Harrietville (Tronoh), Ltd.. . . . . . . . . . . . . . . .. . . . . . 282,100 cu. yd., 252 oz. Au* * Month of May, 1947. NEW SOUTH WALES

The state's largest gold mine and one of the leading mines in the Commonwealth is the New Occidental Gold Mines N.L. at Cobar. The ore in the New Occidental itself is a clean gold ore, but the company has been working two adjoining mines, the New Cobar and the Chesney, producing coppergold ore.


415

The plant includes a cyanide section with a regeneration system for removing copper (see Chap. XIV), while a special flotation section is provided for handling the copper-gold ores. The production for 1945 is shown in Table 87. Mount Morgan. This famous property, which is treating a coppergold ore, milled 67,260 tons in a recent month. The annual production in 1946 was 2795 tons copper and 59,050 oz. gold from ore averaging 4.23 dwt. gold per ton and 1.9 per cent copper. TABLE

87.

GOLD PRODUCTION IN NEW SOUTH WALES,

.,

are production

Tons milled

Value, dwt. per ton

72,662 25,528 38,888

8.85 6.43t 3.59t

CUI

tons

405 881

1945*

Recovery

Au, oz.

28,517 6,691 5,733

Cu, per cent

Au, per cent

98.4 98.9

88.5 81.7 82.0

" Compiled from figures given in E. and lJ1.J., September, 1945, p. 155. t 1.61 per cent Cu. t 2.29 per cent Cu.

QUEENSLAND

One of the important producers of the well-known Cracow gold fields in this state is the Golden Plateau N.L. west of Rockhampton which is cyaniding a quartz ore carrying fine gold. For the year ending June 30, 1946, a total of 82,728 tons was milled with a yield of nearly 50,000 oz. gold. A considerable proportion of the recovery is made on corduroy strakes, NEW ZEALAND

The precious-metal production of New Zealand in 1945 was 372,908 oz. gold and silver. The greater part of this was derived from dredging operations of which the Consolidated Gold, the Chitha River Gold Dredging, and the Kanieri Gold Dredging Companies are the better known of the various companies working the river and placer deposits of this country. FIJI ISLANDS

In a series of six articles from July to December, 1947, in C.E. and M. Rev., H. H. Dunkin describes gold-mining activities in the Fiji Islands. The islands of the western Pacific constitute a remarkably homogeneous major mineral province. A great belt of basic and intermediate igneous rocks extends from the Philippines through Borneo and Celebes to New

416


Guinea and thence to the Solomons, New Caledonia, New Zealand, and Fiji. The andesites and dacites of these islands are associated with characteristically silver-rich tertiary epithermal gold deposits. The Mount Kasi lodes on Vanua Levu are associated with shear zones in a silicified andesite breccia. On the Tavua field the gold deposits occur on and near shears through basalt but are genetically related to younger andesites which exist in the center of a basin but are eroded from its rim, exposing the fractured and mineralized basalt. The gold occurs partly as the native metal and partly as the tellurides sylvanite and hessite. In the Dolphin East lode some of the gold has been proved to be submicroscopically associated with pyrite. The chief gangue mineral is drusy quartz, and some ore consists of propylitized and fractured basalt with tellurides on the fractures in well-developed crystals and with relatively slight silicification. Minor vein minerals occurring in small amounts include marcasite, arsenopyrite, pyrrhotite, sphalerite, bornite, chalcopyrite, covellite, native copper, native tellurium, and tetrahedrite. Although there had been intermittent prospecting for gold in Fiji for more than 60 years and at various times small showings of gold had been discovered in river flats and rock outcrops, nothing of importance was found until the Mount Kasi ore body was located in the Yanawai district of Vanna Levu in the late 1920's. In 1931 a lO-head battery and pilot plant was erected, and on July 8, 1932, the first shipment of 58 oz. of bullion was exported by S.S. Niagara. Later the plant was enlarged, and the company maintained regular production until it closed down in July, 1943. Of over a score of companies operating to greater or less extent on the Tavua field in 1935-1936, only five were active at the end of 1937. However, production from the Emperor, Loloma, and Dolphin mines up to June, 1946, has amounted to more than 1,500,000 tons of ore treated for a yield valued at over £6,750,000 in Fijian currency. Since September, 1944, ore mined by these three companies has been treated in one central plant at Vatukoula" (see Fig. 93), a town of some 3500 population. The ores from the three mines are not blended for treatment but are treated separately. Central Mill at Vatukoula (Type IIIr). The grinding section has two 7- and 6-ft. grate-discharge primary mills, one Marcy and one Ruwolt, in closed circuit with Dorr F 8-ft. by 26-ft. 8-in. classifiers and one 7- and 6-ft. secondary Ruwolt mill in closed circuit with an 8- by 32- by 15-ft.-diameter Dorr F bowl classifier. These grind the ore to 72 per cent through 200 mesh. The Ruwolt mills have been speeded to 24 r.p.m., and recently another similar secondary mill and classifier have been added to increase grinding 29

Hill of gold ("vatu," hill).

417


600-ton fine ore bin

't

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FIG. 93. Flow sheet of the Central Mill at Vatukoula,

418


capacity. When this unit is operating, it is expected that grinding will be improved to 85 per cent minus 200 mesh. Bowl-classifier overflow passes to the 120-ft. thickener. Thickener overflow goes to the Merrill-Crowe plant, and underflow to nine 30-ft. by 11-ft. Ill-in. rake-type agitators in series and the Moore filters. Repulped filter cake is split so that part circulates through gas tower, agitator, and circulating pump, finally overflowing the agitator to be pumped to rejoin the remainder of the pulp stream in the first of a series of four Wallace agitators which are used for conditioning by intense aeration. The gassing section is in duplicate, one unit being used at a time. Gas towers are 10 ft. inside diameter by 42 ft. high, of wooden construction with 17~-in. tongue and groove lining, and having a 6- by 6-in. pulp inlet, 6- by 12-in. pulp discharge, and 4- by 3-ft. gas inlet. The reinforced concrete agitators are 11 by 10 ft. inside diameter by 4 ft. 6 in. deep, with steel paddles. Circulating and agitator overflow pumps are, respectively, 4 and 3 in. Wilfleys with castings of acid-resisting chrome steel. The WallaceDenver agitators are 15 ft. 6 in. by 15 ft. 11 in. with 3-ft. 8-in. four-bladed rubber-covered impellers driven by 20-hp. motors and doing 128 r.p.m. A Wilfley pump elevates conditioned pulp from the Wallace agitators to a 30- by 12-ft. surge tank ahead of the flotation section. It has been found that best flotation results are given by controlling the quantity of pulp exposed to roaster gas so that the mixed gassed and ungassed pulp in the first Wallace agitator contains 0.048 pel' cent sulphur dioxide (determined by titrating a 25-cc sample of solution with 0.40 N iodine). In the Wallace agitators the pulp is subjected to agitation and aeration for 5 hr. and is then pumped to a surge tank from which it gravitates to the flotation conditioners. At this stage the pulp contains 0.003 per cent sulphur dioxide and has a pH of 6.0. Copper sulphate, at the rate of 0.4 lb. per ton, is fed to the pulp in the first Wallace agitator. Butyl xanthate used totals 0.2 lb. per ton, of which half is added with pulp feed to the flotation surge tank and the remainder to the 10- by lO-ft. Denver conditioners ahead of the rougher flotation machines. For frothing, a mixture of equal parts cresylic acid and pine oil is used, a total of 0.07 lb. per ton being added, part to flotation conditioners and the rest to the flotation cells. Only two rougher flotation machines are used, the third being held as a spare, and no cleaning of flotation concentrate is practiced. The two flotation machines operate in parallel, and a finished concentrate is normally taken from the first four Denver cells of each machine, while concentrate floating from the remaining four Denver cells and three Fagergren cells constitutes a middling product which is returned to the flotation feed surge tank. Denver and Fagergren flotation cells have steel tanks, and wear


419

has been fairly rapid owing to the acidity of the pulp. About once 11l 3 months the spare flotation machine is used, while the other machines are overhauled in turn. With correct gassing the telluride and auriferous pyrite float rapidly. There is a moderate amount of latitude permissible in gassing, good results normally being obtained on flotation feed pulp having pH of 6.0 gives best results. The amount of sulphur dioxide required for correct gassing varies with the different ores; laboratory tests show that Emperor and Loloma ores usually require about 25 lb. S02 pel' ton 'while Dolphin ore requires 50 lb. S02 per ton of ore. Half-hourly determinations by the Fijian operators show the sulphur dioxide content of pulp irl the first Wallace agitator and at the feed point of the flotation machines, also the pH of flotation feed, the latter being determined colorimetrically, Conditioning with sulphur dioxide gives good precipitation of dissolved gold. Repulped Moore filter cake normally contains some 0.4 dwt. dissolved gold pel' ton of solids, while flotation residue averages 0.03 dwt, dissolved gold per ton. Final flotation concentrate from the 30- by 8-ft. concentrate thickener passes to a 15- by 8-ft. surge tank and 8- by lO-ft. Oliver filter. Thickener overflow and Oliver filtrate are used at the Moore filter for washing. The filter cake passes to the rotary driers, which are 17 ft. 6 in. long by 4 ft. 17'2' in. diameter inside brickwork and revolve at 37'2' r.p.m, with a slope of % in. pel' ft. These were originally in the Loloma mill, but the central chrome-steel combustion cylinders have been removed, and new feed and discharge hoods fitted. Conveyor belts transfer dry concentrate to a storage bin 19 ft. 6 in. high by 14 ft. diameter whence it is fed to the Edwards roasters. These are both duplex 60-rabble furnaces 130 ft. 9 in. long by 14 ft. 6 in. outside and 11 ft. 6 in. inside width. Rabbles 10 to 20 are water-cooled, and rabble speed is 2.5 r.p.m. Slope is X6 in. pel' ft. Roaster fans are Richardson superlimit load fans, with a capacity 0 15,000 cu. ft. per min. at 120°C. and 6 in. standard water gauge. Fan speed is 960 r.p.m. Cyclone and multiclone dust collectors are installed on flues. The stack is 6 ft. 6 in. diameter by 203 ft. high. A 12-in.-cliameter screw conveyor transfers calcine to a sunken reinforced concrete surge agitator, 15 by 8 ft., where milk of lime is added. Calcine pulp passes to a Wallace agitator and a 5- by 4-ft. ball mill in open circuit. This mill is driven at 30 r.p.m. by a 60-hp. motor, and the calcine is ground to 80 per cent minus 200 mesh. Mill discharge is cyanided for 3 days in three 11-ft. 6-in. by 11-ft. 10-in.-deep standard Dorr agitators and then passes to a 30- by 12-ft. tray thickener. Solution passes to the Merrill-Crewe plant, and underflow rejoins the main pulp stream at the 120-ft. thickener. During the year ending June 5, 1946, 170,481 tons was treated in the com-

420


bined mill for a yield of 86,292 oz. fine gold. The over-all milling cost was 18s 8Ad per ton or £1 16s lId per ounce of gold. Details are given in Table 88. Cassel cyanide is normally added at two points in the mill circuit, viz., to gold solution overflow from a 120-ft. thickener and to a calcine ball-mill feed tame Cyanide strength of gold solution for precipitation is maintained at a minimum of 0.03 per cent sodium cyanide, the pregnant solution over a period averaging 0.04 per cent NaCN. Barren and mill solutions supply all the cyanide used in grinding, thickening, and precyanidation sections of the mill, the free cyanide content falling to 0.020 per cent TABLE

88.

METALLURGICAL RESULTS AT VATUKOULA

Item

Head, dwt. per ton , . Head, sulphur, per cent . . Over-all extraction Flotation concentrate, dwt. per ton . Partly cyanided calcine returned to 120-ft..thickener, dwt. per ton . Over-all residue, dwt. per ton . Cyanide consumption: Precyaniding, lb. cyanide (NaCN) per ton . Over-all, lb. cyanide (NaCN) per ton . Lime consumption: Precyaniding, lb. lime (CaD) per ton . Over-all, lb. lime (CaD) per ton " , .

Emperor

Loloma

Dolphin

8.06 1.45 89.5

30.3 1.65 92.4 240

31. 7 2.95 91.4 320

6 0.85

9 2.27

10.5 2.73

0.81 1.54

0.77 1. 78

1.20 2.49

110

14.2 15.3

13.4 15.0

23.6 25.4

Loss of weight in roasting: 20 to 23 per cent. Roaster discharge contains 0.25 per cent sulphide sulphur and 2 per cent sulphate sulphur. , Reagent consumption in calcine cyanidation: lime, 35 to 40 lb. per ton of calcine; cyanide (NaCN), 24 lb. per ton of calcine.

NaCN in pulp fed to Moore filters. Lime is added as a thick suspension to the primary ball-mill feed and to the calcine repulper. A fairly uniform supply of lime is fed to the primary ball mills to maintain alkalinity of the overflow solution from the 120-ft. thickener at 0.04 per cent CaO. Alkalinity of pulp steadily decreases through the agitation section, and pulp fed to Moore filters averages 0.02 per cent CaO. Lime for calcine cyanidation is added to the repulper to keep solution alkalinity of 0.09 per cent CaO at this point. The calcine is repulped with barren solution and receives up to 24 hr. agitation and aeration before being pumped to the calcine ball-mill feed tank, where cyanide is added to raise the solution to 0.09 per cent NaCN. Pulp discharged from the


421

calcine cyanidation section to rejoin the main plant flow averages 0.04 pel' cent NaCN and 0.04 pel' cent CaO. A flow sheet of the central mill is attached. PHILIPPINE ISLANDS

This is one of the most important gold-mining areas of the world. Its history" goes back many generations. Natives of the Mountain Province were making earrings and other ornaments of gold long before the Spanish explorers penetrated the region. Modern gold mining was initiated only 30 years ago. American prospectors first met with conspicuous success in" the Benguet region, and finally in 1933 the increased price of gold encouraged the development of low-grade properties previously regarded as unprofitable. The principal gold-bearing districts in the Philippines are (1) Mountain Province, (2) Camarines Norte, (3) Suraigao, and (4) Masbate. The first two are located in the island of Luzon, the distinction between them being made because ore from the lodes of the Mountain Province (and Masbate) is simple in character and well adapted to free milling treatment while that from Carmarines Norte and Suragao is highly refractory, containing substantial amounts of copper and other metals, which necessitate the use of flotation and smelting methods. Practically everyone of the 33 mills operating in 1941, with a production of 1,130,000 oz. gold, were partially or totally destroyed during the war, and only a few have yet been able to resume operation. Table 89 shows the 1940 figures for the principal producers, with notes as to present status. The first mill to start up after the war was the Atok Big Wedge. Its 400-ton mill was put into operation in March, 1947, and at reduced capacity milled 19,201 tons of 0.52-oz. ore in 9~~ months. 'One of the most important postwar developments in the Islands is the establishment of a central mill" at Baguio to treat ore from Benguet, Balatoc, and Cal Herr mines, the plants of which were destroyed during the war. The total reserves of the first two properties are estimated at 4,000,000 tons with a gross value of over $31,200,000. The mill is rated at 1200 tons per day, later to be expanded to 4000 tons pel' day. The Suragao Mother Lode has started milling a small tonnage, some of which averages 0.88 oz. gold, and the Suragao Consolidated is milling around 5000 tons per month of a low-grade are from its open pit. Half 30

Report of the National Development Co., under the supervision of the H.

E. Beyster Corporation, consulting engineers, Oct. 28, 1947. 31 Article by W. F. Boericke, E. and M.J., March, 1948. The author is indebted to Boericke for much of the information regarding the present status of mining in the Philippines.

422


the recovered values were shipped to the United States as table and flotation concentrates, the rest being recovered as bullion that is sold locally. The Lepanto Consolidated, generally classed as a copper mine though containing good gold values, was building a 500-ton flotation plant scheduled to go into operation in March, 1948. TABLE

89.

District

Luzon

Masbate

PARTIAL LIST OF PHILIPPINE GOLD PRODUCERS*

1940tons milled

Company

Benguet Consolidated Balatoc Itogan Atok Big Wedge

467,130 738,716 327,062 75,642

Baguio Gold Demonstration Antamoc Goldfields Lepanto:j:

142,545 44,360 193,966

Suyoe Antipolo United Paracale Paracale Gumas San Mauricio North Camarines

79,152 5,736 126,321 71,922 182,716 63,944

... - ...

Av. value, t oz. per ton

Type of mill and present status

0.40 }Cyanide, 1200-ton postwar 0.38 plant now in operation. 0.23 Cyanide, destroyed. 0.40 Cyanide, 400-ton plant now operating. 0.22 Cyanide, destroyed. 0.16 Cyanide, destroyed. 0.19 Cyanide, destroyed. .. .. Cyanide, 500-ton flotation plant now being built. 0.36 Cyanide, destroyed. 0.27 Cyanide. 0.37 Flotation, destroyed. 0.38 Destroyed. 0.47 Flotati on, destroyed. 0.34 Destroyed.

Masbate Consolidated 1,078,573 0.085 Cyanide, destroyed. Capsay 43,183 0.35 Cyanide, destroyed. LX.L. 144,868 0.30 Cyanide, destroyed.

Mindinao (Suragao) Mindinao Mother Lode Suragao Consolidated Davao

* Prewar-tonnage figures. t Chamber of Mines of the

68,487 0.45 }Flotation, limited tonnage 117,012 0.27 being milled during reconstruction program. 4,334 0.31 Destroyed.

Philippines.

The figures in pesos per ton are recal-

culated at 70 pesos per ounce. :j: Copper-gold ore.

In addition to the lode gold mines, there were five placer operations listed in 1941, with a total output of about 1 million dollars from the 4H million cubic yards handled. Several of the smaller operations have been resumed, but no attempt has yet been made to rehabilitate the large dredges in the Paracale field.


423

The total gold production of the Philippines for 1947 was about 100,000 oz., and for 1948 incomplete reports give a figure of 200,000 oz., which is less than one-fifth the prewar production and is indicative of the extent of the damage to the industry during the war years. Atok-Big Wedge Gold Mining Co. (Type IIa). This well-designed mill, the first to be rehabilitated since the war, is known to be one of the most metallurgically efficient plants on the island. The ore carries calcite and quartz with some pyrites, sphalerite, and chalcopyrite. The free gold occurs in a very finely divided state.: The mill has a capacity of 450 tons per day, and the feed averages 0.50 oz. ., gold and 0.28 oz. silver pel' ton. The ore is crushed in two stages to X in. and is then passed to a twostage grinding circuit consisting of Marcy and Traylor grate-discharge ball mills closed-circuited with 16-ft. by 18-ft. 4-in. Dorr classifiers in the primary stage and Traylor ball mills closed-circuited with Dorr bowl classifiers in the secondary stage. Air lifts are used to elevate the pulp from the mill discharge to the classifiers. The pulp ground to 50 to 55 per cent minus 200 mesh and overflowing the, bowls at 18 per cent solids, is split between two parallel circuits which present several unusual features. Agitation is carried out without thickening in two sets of four 18- by 18-ft. Dorr agitators. Then stage washing (rather than C.C.D.) is done in each circuit is a series of three 3Ll-ft. 6-in. by ll-ft. 6-in. Dorr thickeners, adding mill solution to the second thickener and barren solution to the third. Finally, pregnant solution is overflowed to the pregnant-solution tank from each thickener. The last thickener underflow passes to an eight-section Moore filter plant operated in two parallel units. The filtrates all pass to the mill solution system, and the cake is given three barren washes followed by a final water wash before being discharged to the tailings line. Clarification is carried out in two 15-leaf tanks, and the solution precipitated by the Merill-Crowe system using bag precipitation boxes. Precipitate is melted to bullion on the property. The tailings average 0.022 oz. gold per ton, which is equivalent to a 95.6 per cent recovery. The power consumption is distributed as shown in Table 90. ASIA

Russia. There are no published figures on the exact gold production in the U.S.S.R., but the estimated world total of 27 million ounces in 1946 credits Russia with a figure of 6 million ounces.f This figure agrees reasonably well with the 225-million-dollar output (6,!/z million ounces) 32

Minerals Year Book, 1946.

H"l-..? H"~ ---Orepulp ----Wolsr

0-------- Mill sotatton

- - - - - PfBr;nonf so/aI/on

_._-- Bamm solution

o

~ ~ tj

~

o

z p.

ztj o

o

Z

o l=J Z t-3 ~

~

H

o

Z

o "'l

o

~

l=J

U1

rlmnBllJr!ver lIDa.olGol-/T

@3C"molslllre

aD 0.001

041

FIG. 94. Plan of the Atok-Big Wedge mill showing ore pulp, water, and solution flow.


425

estimated by Vladimir Petrov in his article on Russian gold production in the magazine section of the Herald Tribune. for July 11, 1948. 'Petrov believes that the Siberian gold fields on the Kolyma River yield not less than three-quarters of the total Russian output.: The gold mined in this extensive area is recovered by gravity methods, and evidently mining communities of considerable size have been developed in recent years. Rich mines such as the Ridder and Sokolni were also worked at one time in the Altai Mountain district, and gold and platinum are known to have been recovered by dredging operations in the Ural Mountains. It is probable that both of these areas are being exploited today, since the Soviets seem to regard gold production of special importance in their present economy. TABLE

90.

POWER DISTRIBUTION AT ATOK-BIG WEDGE

Hp.-hr. per ton milled

Operation

Crushing, , .. , . , Grinding and classification.. Cyaniding, , Filtering, , . , , .. , ,., , Precipitation , Lighting., , , Water supply .. , . , Miscellaneous", TotaL., ,


1. 719 12.09 59.42 8.449 2.197 . 15.45 0,654 . 4.60 . 0.796 5.60 0,282 . 1.98 . 0,82 0.117 0,006 . 0.04 1-----1----. 14.220 100.00

, .

India. The most important gold-mining area in India is that of the Kolar gold fields in the state of Mysore, Here the gold occurs in quartz reefs and is free milling. The treatment scheme is all-sliming and includes blanket concentration followed by cyanidation. Through J ohn Taylor and Sons, managers, and a subcommittee of the chief metallurgists of the group of four plants operating at Kolar, Mysore State, India, it is possible to give current treatment practice. (The detailed and excellent paper of R. H. Kendall and A: F. Hosking in Trims. 34, I.JJ!I. and lJ1., Part 2, 1924, or in Proe. Empire 111. and lJ1. Conqr. (London), Part 2, 1925, may be remembered by metallurgists; since then amalgamation in the stamp- and tube-mill circuits, airlifts, and sand treatment have been eliminated.) The Kolar group of mines is a remarkable one for depth attained (between 7000 and 9000 ft.), ore persistence, ore reserves, ore production (34,400,000 tons), gold production (21,000,000 fine oz.), and dividends (£27,900,000 ore, $112,000,000) during more than 60 years to the end of 1947. The ore of the four mines comes from the Champion lode or its

426


branches and is free milling. The quartz is hard and abrasive, and free gold is often seen. The average gold content in 1947 was 6.82 dwt. Galena, pyrrhotite, pyrite, and mispickel are accessory minerals but rarely exceed 1 per cent. At Nundydroog parallel western lodes containing up to 13 per cent of these minerals are also being worked. The general practice is as follows: The ore is delivered to sorting and crushing stations and passed over grizzlies set at 17~ in.; the coarse material passes to picking belts or tables for removal of waste rock and tube-mill pebbles and is then fed to jaw or gyratory crushers set at 17~ in. Both products are trammed to the stamp-mill bins, being weighed en route. The waste rock picked out amounts to about 10 per cent, some of which serves as tube-mill pebbles. During 1947 an average of 170 head of stamps was in use. Screens of 2 or 3 mesh are general; thus, the minus 1H-in. feed is reduced to minus 7~ in. Blanket tables and machines are used to catch gold and sulphides after both stamps and tube mills; Nundydroog relies on posttube-mill concentration only. The combined pulp from the stamps and tube mills is pumped to a distributor. From this, the pulp flows to primary and secondary classifying cones; the underflows are fed to the tube mills, while the overflow from the secondary cones is delivered to the slime collectors. All ore is ground to a final product of 80 to 88 per cent through 200 mesh. Slime treatment consists of collecting and thickening in cone-bottom tanks, pumping the thickened pulp into either mechanical or Brown agitators for contact with cyanide and thence into a stock ore surge tank. Filtration is carried out in Butters-type filter installations, the treatment cycles of which occupy approximately 100 min. Residue is pulped and pumped to the dump. Gold is precipitated from cyanide solutions on zinc shavings, which are cut locally. Precipitation is practically complete. Mill products and solutions are sampled at. all stages of operations. During 1947, 508,217 tons of ore was treated in the four plants: Mysore, Champion Reef, Ooregum, and Nundydroog. The ore averaged 6.82 dwt. per ton, and the residue 3.26 grains, equivalent to an extraction of 98.01 per cent. The cost was 100.5d or $1.66, per ton. The consumption of cyanide averaged 0.543 lb. per ton of ore. Burma;" Reports from the Bawdwin mine in Burma, which was the leading silver producer of Asia in prewar years, indicate that substantial output would require several years for reestablishing transportation facilities, recruiting a labor force, and rehabilitating mine and surface plant. In 1939 the Burma silver production was close to 7 million ounces. as

Data obtained from Minerals Year Book, 1945.


427

China and Manchuria. 33 These countries reported a gold production in 1940 totaling nearly 600,000 oz., but as a result of the war and the unsettled political conditions since, it is probable that the industry today is largely inactive. No production figures are available. Japan and Korea (Chosen);" Before the World War both of these. countries were large producers of gold and silver. In 1940 the output of each totaled about 900,000 oz. gold with 10 million ounces silver (Japan) and 272' million ounces silver (Korea). The industry was largely crippled, however, by the effects of the war, and today no production figures or technical data as to methods used are available. oJ New Guinea. The Bololo Gold Dredging CO.34 had a fleet of eight all-electric bucket dredges with a total capacity of about 2 million cubic yards per month working on its leases along the Bololo River. All operations ceased, however, early in 1942, and much equipment was subsequently destroyed as a result of the war. The dredges are now being put back into operation, but lack of power-plant equipment is holding up production. Between the years 1932, when this area was opened up, and 1942, 119 million cubic yards were handled with the production of 1,297,416 oz. gold and 5'Z5,726 oz. silver. 34

C.M. and M. Rev., January, 1947.

CHAPTER XVI

'Treatment Of Silver Ores The previous chapters have been elevoteel to the treatment of golel anel silver ores in which the recovery of silver, because of the relatively small amount present, is not orelinarily of economic importance. There are, however, certain mining areas where the recovery of the high silver values is 01' has been the principal metallurgical problem. The present chapter diecusses some of the current silver-treatment plants anel also reviews briefly some of the olde: practices in important silver-mining areas since closeel elown. The greater part of the worlel's prcductior; of silver is derioed from the refining of the base metals, particularly leael ores, anel complex ores of leael, copper, antimony, anel zinc. J110st of these ores are concentrateel by flotation rnethoels, anel the concentrates smelted: There are silver ores, however, where the base-metal content is too low to just~fy the above conventional form of treatment, anel cya'11ielation offer» the most economic recovery method.

Table 91 gives some of the more important cyanidation and flotation data obtained in laboratory tests on relatively pure samples of silver minerals. NATIVE SILVER

Cobalt District, Ontario, Canada. A very complete history and description of the mines, mills, and metallurgical treatment of the Cobalt, Ontario, silver ores have been written by J. J. Denny, Fraser D. Reed, and R. H. Hutchinson and published in 21st Ann. Rept., Onto Dept. lI1ines, 1922. Today not one of these mills is in operation. It is of interest, however, to note that between the years 1904 and 1919 some 30 operating mines shipped 153,874 tons of high-grade ore averaging 1000 oz. silver per ton and 122,130 tons of concentrate averaging 664 oz. silver per ton, which, together with bullion shipments, represented a total output of 183 million dollars. The ores of the Cobalt area were remarkable for their high content of silver and for the complex assemblage of minerals found in the veins and enclosing rock. Of the silver-bearing minerals, native silver was of outstanding importance, as fully 97 per cent of the values occurred in this form. It was found in masses ranging from large slabs to the finest, filmy leaf. Other minerals included cobalt and nickel in the form of arsenides, 428

TREATMENT OF SILVER ORES

429

sulphides, antimonides, and various combinations of these, associated ami often intimately mixed with a number of base-metal compounds.' A variety of methods were used for treating high-grade ore and concentrates, including the amalgamation and cyanide process, the hypochlorite-cyanidation process, the sulphuric acid-cyanidation process, and' chloridizing roasting, while the lower grade material was treated by a combination of gravity concentration and flotation or cyanidation. SILVER SULPHIDES

Shafter, Tex. (Presidio). Between 1883 and 1942 when it was closed down, the Presidio mine of The American Metal Company, Shafter, Tex., produced more than 55 million ounces silver from an ore averaging 10 to 20 oz. silver pel' ton. An 85 pel' cent recovery was made in the 400-ton gravity concentration and C.C.D. cyanide plant, which is well described by D. E. Stem in Mining J., Apr. 15, 1941. The ore was oxidized and siliceous, the principal constituent of the gangue being quartz with some calcite. The silver minerals contained in the ore were principally argentite and cerargyrite, the former predominating. The lead minerals, all of which were argentiferous, were chiefly cerussite and galena, with occasionally a little anglesite. The gold was free, but most of the ore contained merely a trace. The milling scheme included tabling at 10 mesh followed by regrinding to 80 to 90 per cent minus 200 mesh and cyanidation by Pachuca agitation. The solutions were maintained at 2.8 to 3.2 lb. pel' ton NaCN, and the reagent consumption was 6 lb. lime, 2.5 lb. NaCN, and 0.25 lb. zinc dust pel' ton of ore. Table concentrates averaged about 426 oz. silver per ton and carried 52.5 per cent lead. The zinc precipitate analyzed 20,000 oz. silver and 5.60 oz. gold pel' ton and carried 1.02 per cent zinc and 24.1 pel' cent lead. Both products were shipped to Carteret, N. J. Pachuca, Mexico. Silver in the Pachuca district, state of Hidalgo, Mexico, occurs chiefly as argentite. A' part composite analysis of ore treated at the present time is given in Table 92. Flotation has been given thorough trials but has not succeeded in equaling the economic results of cyanidation, according to R. R. Bryan and M. H.Kuryla in Trans. 112, A.I.M.E., 1934. The Loreto plant of Compafiia de Real del Monte y Pachuca has a daily capacity of 3800 tons and is the largest silver-cyaniding works in existence. No concentration is done. The property was purchased by the Mexican government from the United States Smelting, Refining and Mining Co. in September of 1948. For several years past the tonnage and grade of ore has been dropping and now stands at about 100,000 tons pel' month, assaying 300 grams silver

430


91.

CYANIDATION AND FLOTATION

Mineral composition and cynnidntion extruetions Argentitet Ag,S

Cerargyrite AgCI

Bromyri te Embolite

silver glance

87.1% Ag

horn silver

75.3% Ag

87% in 72 hr. 93% with Nu,O, 97 to 100% All dissolved in NaCN with good extractions.

AgBr Ag(Br, CI)

Native silver (when finely divided: wben coarse dissOlVeS) very slowly) Proustite Ag,AsS, Light ruby silver

42.5% warm 72 hr. 91% roast, changing solution.

05.4% Ag

Pyrargyrite Ag,SbS, dark ruby silver 59.8% Ag All do not dissolve so readily.

07% warm 72 hr. 88% roast, changing solution.

Stephunite Ag,SbS.

08.5% Ag

90% in 72 hr. Not improved by roak

Polvbusitej

75.0% Ag

80% warm 72 hr. 90% after roast.

AgDSbS,

Argentiferous galena PbS Ag in solid solution Stetefeldtite Ag, CuO, FeO, Sb,O,S

Tetrahedritell

99% in 24 hr. exeeption.)

CUBSb,S, gray copper. Ag replaces Cu

Sphalerite ZnS

zinc blende, Ag in solid solution

Ag tied up with these minerals usually cyanided with great difficulty.f

(An

25% standard. 70% after roust.

Silver jarosite Ag,Fe,(OH),,(SO.). like plumbojarosite Ag in manganese minerals

Ag leached by H20, in acidic solution or reduction roasting.

• Abstracted from: "Oxygen as an Aid in the Dissolution of Silver by Cyanide," R.I. 3004, U.S.B. of M. and "Flotation of Silver Minerals," R.I. 3430, U.S.B. of M. tAg dissolves according to the reaction: AgCl 2NaCN = NaAg(CNJ, NaCl Where argentite was intimately mixed with pyrite, sphalerite, and gangue, roasting for 1 hr. up to 400'C Wasnecessary for a 75 per cent extraction. Time is the important factor for argentite when pure. If temperature of roast exceeds 600°0 an insoluble silver silicate is formed. § In the case of polybnsite, silver 'may be partially replaced by copper and antimony partially replaced by arsenic. II Not all samples of tetrahedrite areas refractory as the one tested. Some yield up to 83per cent extraction. ~ In general the treatment includes 1. Fine grinding. 2. High alkalinity. 3. Lead and mercury salts to precipitate alkaline sulphides Tests sbowed that 1. Oxygen was an aid in dissolving silver minerals. 2. Warm solutions are an advantage. 3. A low-temperature roast gives the best results of all.

+

+

+

431

TREATMENT OF SILVER ORES CHARACTERISTICS OF SILVER MINERALS*

Rntio of concentration

recovery

Recovery lowered by iron oxides.

107.4:1

98.5

Normal fictubion. Lime has little effect. Concentrate grade lowered by slimes but improved by starch.

25.5: 1

98.8

Lime deleterious. Grade hut not recovery lowered by talcose material. Starch cannot be used.

12.4: 1

94.5

t Lime deleterious. Both recovery' and grade lowered by slimes, but starch

68.7: 1

97.0

Lime deleterious. Talcose slimes lowered grade only, but starch can be used to correct this.

87.5: 1

94.4

Lime has little effect. Talcose slimes lowered grade only! but starch can be used to correct this.

07.5:1

98.7

Lime has little effect. Taleose slimes lowered grade only, but starch can be used to correct this.

243.0: 1

99.1

Flotation characteristics Normal flotation. Lime has little effect. Starch helps concentrate grade.

Best

.

oJ

cannot be used.

t,Pyrargyrite is very sensitive to changes in flotation conditions. Sodium sulpltide was definitely "armful in all cases. The pure minerals were mixed with Sell sand and floated with and without the further addition of talcose and iron oxide slimes. Following dry crushing, the mixture was ground in a pebble mill to minus 65 mesh. Each sample was then floated in a Denver Sub-A laboratory cell for 10 min. at 22 per cent solids. The reagents used were Amyl xanthate •0.5-1.0 lb. per ton Aerofloat 15 0.1-0.2 CresyJic acid , " 0.05-0.2 The feed in the case of the synthetic mixtures averaged 15 to 70 oz. per ton. And where natural ores were tested, slimes were also found to have a noticeable influence on results. In a general way the results obtained with natural ores tended to confirm those obtained with the synthetic mixtures.

432


and 3 grams gold per ton. The higher ratio gold than previously is due to the discovery some 5 years ago of a new vein carrying about 10 grams gold per ton. Some details of 1948 operations of the Loreto follow (see Fig. 95): Crushing. Mine are of a maximum size of 12 in. is reduced to ~B in. by one gyratory and two cone crushers, between which are grizzlies and vibratmg screens. Grinding. Two-stage grinding in cyanide solution is practiced. For primary grinding, 8- by 6-ft. Marcy grate mills and 6- by 12-ft. trunnion Traylor ball mills are used in closed circuit with 6- by 22-ft. Dorr classifiers, which are the only type operated in Pachuca. Nearly 80 per cent of the feed to these mills is coarser than 3 mesh and up to 1 in. The classifier overflow is 64 per cent solids. For secondary grinding, 6- by lO-ft. Traylor mills of the trunnion type and 5- by lO-ft. trunnion mills of local make are used in closed circuit with an 8- by 22- and a 6- by 22-ft. classifier, respecTABLE

92.

CHEMICAL ANALYSIS OF PACHUCA ORE

Constituent

Per cent

Constituent

Per cent

Silica ........................... Alumina. ...................... Iron. .. .................... . . .. Lead .. ....... - ................ . Zinc. ..........................

71.6 8.7 2.9 0.05 0.5

Copper ...................... Manganese. . . . . . . . . . . . . . . . . . Sulphur . . . . . . . . . . . . . . . . . . . . . Lime and magnesia .... . .. . . Sodium and potassium oxide.

0.05 1.6 0.7 3.8 1.9

tively. The classifier overflow contains 20 per cent solids; a sieve test of the final product shows on 48 mesh 2.20 per cent; on 65, 7.90; on 100, 9.17; on 150, 13.13; on 200, 8.09; and through 200, 59.51 per cent. Thickening. Ten Dorr thickeners, 48% by 15~~ It., yield a pulp of 45 per cent solids. Between 1934 and 1945 the are became more difficult to settle and underflows dropped' to as low as 30 per cent solids. The trouble was largely overcome by removing about 1000 tons per day of plant solution and replacing with fresh cyanide solution. The solution r:emoved is plant barren solution, which is first passed through the regeneration plant to recover its cyanide, silver, and gold content and then discarded with the tails. Agitating. Eighteen Pachuca tanks, 15 by 60 ft., and 32 "flat" tanles,' 20 and 24 by 30 It., do the agitating. The latter is a tank equipped with a Dorr-thickener mechanism and air jets. Air at 35 lb. pressure is used in the Paohucas and at 18 lb. in the flat tanks. Agitation proceeds for 73 and 70 hr., respectively. Aero-brand cyanide is dissolved in barren solution to make a strong solu-


433

tion, and this is added to the agitators to bring the strength to 0.17 per cent NaCN. Litharge is added in the dissolving tank to 'eliminate soluble sulphides. Cyanide consumption, excluding regeneration, amounts to 1.62 kilograms pel' ton ore. Lime consumption is 9.0 kilograms. Filtration. Butters tanks, each with 187 leaves, 67 by 117 in., do the filtering. Each tank averages 11 cycles of 128 min. each day, and each cycle is divided into 26 min. for caking, 38 min. for barren wash, 15 min. for water wash to mill, and 20 min. for water wash to regeneration, the remaining 29 min. being required for filling transfers, discharging, etc. A vacuum of 18 in. is maintained. Average cake is 7-& in. thick. Clarification and Precipitation. Solution from the filters is clarified in 12 Sweetland presses, which can handle 2)'4 tons per day per square foot of surface. They are discharged twice and cleaned once each day, and leaves are acid-treated every 10 days. The Merrill-Crewe system of zinc-dust precipitation is used. Centrifugal pumps force the solution through the presses. Zinc consumption is 170 grams pel' ton of ore. The dried precipitate assays 83 pel' cent silver and 0.46 per cent gold, also 0.25 per cent selenium and some other metals. Melting and Refining. Precipitate is melted to bullion in the usual manner, granulated borax and bottle glass being used, in an oil-fired reverberatory furnace of 15 tons' capacity. The temperature is raised to 1050°C., and slag is skimmed off. Air is then blown in, and the slag is skimmed for 60 hr. Then the metal is tapped into a continuous anodecasting machine. Anodes weigh 10 kilograms, and a furnace charge makes 2000 of them in 5 hr. of casting. The bullion is increased in fineness from 950 to 993; copper is the principal impurity remaining. The anodes are next parted in 200 Thum-type electrolytic cells, and the resultant silver is 999 plus fine. The gold mud is reduced to anodes, which are treated in Wohlwill cells, giving gold 999.8 fine. Extraction. The current extraction (1948) now averages 85 per cent of the silver and 90 per cent of the gold contained in the are. Cyanide Regeneration. The plant for the regeneration of 3800 tons of cyanide solution per day is described in Chap. XIV. Treatment of Silver Ores at Tonopah, Nev. The old milling practice used at Tonopah, Nev., well-known silver district, is of interest to metallurgists today because of certain special treatment features discussed below. 'Writing in 1912, von Bernewitz ' says: In Tonopah there are five mills: the Belmont, Extension, MaeNamaru, Montana, and West End, while at Millers, 12 miles north, are the Belmont and Tonopah mills, ore being shipped to these at a cost of 70 cents per ton. In nearly every case gyratory 1

Cyanide Practice 1910 to 1913, p. 506-509.

,.... CD ~

Miscellaneous operations (Truckhaulage)

SanJuan Pachuca shaft

o

>-1

eI

...

CRUSHING i

PRIMARY GRINDING

T

,

uocrosne: are storage-3,000 metric

tons

Milling solution

H

o

Traylor 20 "gyratory crusher- 31--/'product

Z'1

2-Symons 4 'x 8 'rod deck screens

~

2-Symons 5~/sld cane crushers 3,I/producl

o

i>

o o

Sal7)0ling and weighing Crushed are slarage-2O,OOO metric Ions

z

Lime emulsion

LIME EMULSION PLANT

>-3

~ H

o

Z

o

OJ ~

lJYUff'Ut;;:

Circuit No. 30 aiam tank I ?-24'deep a' 3-20'deep 25%lmilllannage

3

~

o

AGITATION

?-24'deep 3-20aeep 12-2o'deep

oi:'J ;> >-3

THICKENING

2

o;;>-3

4" Grizzly

4 '-6'x 12' Traylor ball m/~/s 4~6'x12' Oorr classifiers 50% mililannage 9-5'x10'ball mills 9-6,r22'Oorr classifiers 36 %mill lannage 10-50'x15'Oarr thickeners

SECONDARY GRINDING

Z H

25 %Imililannage

25 % mililannage

I

Mixer

CYANIDE SOWTION PLANT

i:'J

U1

FILTRATION Thickener overflow solution

CYANIDE REGENERATION

4-6'x12Crowe vacuumtanks ~o->

CLARIFiCATION Zinc dust feeder

o-j

!=:I

t=J

t>

H

""§Z o-j

Q

Gold-Silver- Copper precipitate to smelter

Barren 'sDiUiiOn Mif/ i/~~

'\Z:/

o";I

"11111"1t!!1I1111t11t1l1t7_44plote Merrill

p,rccipilale filterpresses PRECIPITATION Borax and -'~ scrap gloss ~ 2-F?everberatory furnace

:-±-.,

U1 H

t"'

<1 t=J

!=:I

Mill to/Is

-_ Cos/ing machine

• Anodes ;0 parting plant ~

MELTING

o

!=:I t=J

Ul

FIG. 95. Flow sheet of the Loreto mill at Pachuca, Mexico.

H"(X>

I:.H

CYANIDA'l'ION AND CONCENTRATION OF ORES

crushers are used for breaking ore as it comes from the mines, the procedure being to crush first in a large crusher up to the No. 71tl type E: Gates size and pass through revol ving trammels, the oversize being aguin reduced in No. 3 size gyratories, tho final product for the stamps being about 11 :1 in. Sorting is done at the Belmont and J\bcNamam mills, at the former ana pan conveyor from which 15 per cent is rejected, and a.t the latter on II 30-in. rubber belt from which 6 per cent is sorted out. From the crushing depnrtment, the arc is taken to mill bins by 20-i11. belt conveyors, or bucket elevators, and distributed by the usual automatic devices. There is no amalgamation at Tonopah, nor is it necessary on this class of ore. Crushing is done in weak and warm (from 50 to 80°F.) cyanide solutions, so the are is in contact with solution from the stamps to filtration. This is necessary as well [IS the heating, which, although somewhat expensive, quickens the solution and accelerates the dissolving action. Solutions are usually heated to about 950 and in one case to 1200 by live steam introduced in the agitators. The practice of using hot solutions is briefly as follows: At the new Belmont mill the temperature at the stamps is from 60 to 70°F., and at the Puchuoa agitators exhaust steam from the mill air compressor is fed in, increusing it from 90 to 1000 • In the ill. and S. Press of Jun. 27, 1912, A. H. Jones, metallurgist at this plant, gave some valuable duta on this subject. On an are carrying 0.05 oz. gold and 18.2 oz. silver per ton, 60 hr. ngitution with both 60 and 900 solutions, the tailing averaged 0.0175 and 3.45 and 0.0125 and 1.90 oz., respectively. Tests on 48 and 69 hr. at similar temperatures gave as marked results. Besides the effect on extraction, the hot solutions flowing through the mill kept the whole. place at a good working temperature. At the Montana-Tonopah, are is crushed in 50 to 600 solution, which is increased to 110 at the Hendryx agitators by live steam. It is found also that the heat [Lids settling. There is a marked decrease in extraction without hot solutions. Tonopah ores carry as much as 3 per cent pyrite, but concentration is not always employed, it being done only at the Belmont, Montanu, Tonopah, and West End. It would seem that, if the grade of the are and percentage of mineral are not too high, tables are not necessary, and this varies from time to time in the various plants. At any rutc, a very close saving is not attempted. The Extension Company dispensed with their Deister tables, selling them to the West End. The Belmont, Montann, and Tonopah use Wilfley tables. Concentmte is collected, steam dried in large trays, sacked, and shipped to smelters. Freight and treatment cost nearly $70 per ton. All-sliming is the standard method, with the exception of the Tonopah mill at Millers, where three products arc made : concentrate, sand, and slime. At this plant reduction is by stamps and Chilean and Huntington mills, while at Tonopah the procedure is as follows: The pulp from the stamps is fed into Dorr duplex classifiers making 12 strokes per minute, from which slime overflows and coarse material is fed into tube mills by means of a special feeder. Discharge from these is elevated to the Dorr classifier», where II further classification takes place, followed by further grinding in the tube mill, and so on. Vari ous types of thickeners or dewaterers are in use, the practice being to allow the clear solution to overflow and decant off as much as possible for battery storuge, When it gets too high in gold content, it is decanted to the tank for precipitation. As at many other mining centers there is quite a difference of opinion regarding the efficiency of agitators, the Trent being used at the MaeN nmaru., Montana, and 'West End; the Hendryx at the Montana; Pachuca tanks llt the new Belmont mill; and ordinary mechanical agitators and air lifts at the Belmont and Tonopah at Millers, 0


437

these being in series at the Belmont plant. Centrifugal PUIl1PS' and air at about 20-lb. pressure are used for the Trent system, and better results are obtained if pulp is drawn off near the top of a full vat and pumped through the arms as usual. Agitation proceeds for upward of ,18 hr. At the new Belmont mill, slime is first agitated in six Pachuca tanks, and from these it is elevated to Dorr thickeners by an air lift, prior to going to another set of six Pachucas, making a total of 48 hr. agitation, th~ . idea being to get rid of as much valuable solution as possible before sending slime to the filter plant. Cyanide and lead acetate are added to the agitators, the former being from 2 to 5 lb. solution, while regular addition of the acetate is found nenessnry at all mills. Lime is usually slaked and added to the tube-mill feed. Consumption of chemicals at the Extension is as follows: Lead acetate Cyanide. . . . Lime. .

.' 0.9 lb. per ton . . . . . . . . . . . . . . . . . . . . . . . . . . . .. 2.5 . 3.5

Agitated slime is drawn off to stock tanks, which serve the purpose of storage from agitators and excess from filter plants. The latter have little of special note about them, being of the ordinary stationary leaf type which hns been described so often in technical papers. Zinc-dust precipitation is used at the new Belmont and Montana mi11s, and zinc shavings at the Belmont, Extension, MacNamara, Tonopah, and West End. Mothods of dealing with precipitate vary somewhat. At the new Belmont precipitate is dried, mixed with 5 per cent borax, and smelted in double-cornpurtment, oil-fired Rockwell furnaces lined with carborundum, kaolin, and water glass. At the Extension it is dried, fluxed, and smelted in oil-fired Steele-Harvey t il tiug furnaces which contain a No. 250 graphi te crucible, while at the Tonopah mill the fine aincshaving precipitate is incompletely dried, mixed with crude borax which swells up through the mass, and then smelted in six coke-fired tilting furnaces. Crucibles last from 90 to 130 lu-, and are turned once. Tonopah bullion will average 950 fine in silver and a trifle over 10 in gold and is sampled by being bored at opposite corners of top and bottom bars. The bullion is shipped by freight like any other 111.erchandise.

The Montana Tonopah closed its 500-ton mill in 1923, and thereafter no operating mill in the district employed concentration. The average gold extraction in this district was 94 pel' cent, and the average silver extraction 92 per cent. The Sunshine mill operated by the Sunshine Mining Company is situated 6 miles from Kellogg, Coeur d'Alene district, Idaho. The mine is the largest silver producer in the United States. In 1937 its output was 12,147,719 oz. silver and 2,784,289 lb. copper. In 1947, on a curtailed basi" due to labor shortage, it produced 5,034,160 oz. silver, 1,249,555 lb. copper, and 5,881,796 lb. lead. The 1200-ton mill employs the straight flotation flow sheet shown in Fig. 96, the concentrate being shipped to a lead smelter. In 1947 the ore averaged 44.5 oz. silver (associated with galena and tetrahedrite, Cu SSb 2S 7) , 0.55 per cent copper, and 2.61 per

CYANIDATION AND

CONCj~NT11A'l'ION

OF 011ES

•

cent lead. The recoveries were 98.::i3 pel' cent of the silver and 98.01 per cent of the lead. Milling costs were 96 cents pel' ton. The flotation reagents used are 0.13 lb. pel' ton butyl xanthate and 0.73 lb. pel' ton frother. The frother is n, mixture of 1 part Barrett No.4 with 3 parts methyl amyl alcohol. Average concentrate analysis for 1917 is shown in Table 93. The New York and Honduras Rosario Mining Company operates two mills in Honduras, the Rosario and Mochito mills, and the EI Dorado mill in EI Salvador. The following information in regard to the latest practice at these mills was supplied to the author through the courtesy of the president of the company, YV. A. Prendergast. Rosario Mill (Type IIa). The mill treats 550 tons daily of an ore carrying 13.2:'5 oz. pel' ton silver, 0.071 oz. pel' ton gold, O.i:i pel' cent zinc, 0.5 pel' cent lead, and 2.0 pel' cent manganese. Primary crushing is carried out in two gyratory crushers making; a 2-in. product, 350 tons of which is crushed in twenty 1800-lb. stamps and 200 TABLE

93.

AVERAGE CONCEN'.rRATE ANALYSIS AT SUNSHINE

Silver (Ag ) Lead (Pb ) Copper (Cu ) Zinc (Zn) Bismuth (Bi ) Antimony (8b) Arseni c (As)

. 376.55 oz. . 22.00 pel' cent . '1.67 per cent 2.46 pel' cent . . 0.045 per cent . '1.2-[ per cent 2.27 per cent .

tons in a 6- by 5-ft. Allis-Chalmers ball mill charged with 5-in. alloy-steel balls. This mill is in closed circuit with a 5- by 25-ft.6-in. DSFXM Don classifier overflowing a 35-mesh product. The stamp milling is carried out in cyanide solution (3 lb. KCN pel' ton of solution). The product passing the~·i-in. battery screens is dewatered in two 6- by 20-ft. Don DSC classifiers, the overflow going to thickeners and the underflow to two 5- by 9-ft. ball mills in closed circuit with two 6- by 18-ft. DSC Don classifiers, also overflowing a 35-mesh product. A rationed hall charge of 70 per cent 3-in. and 30 pel' cent 4-in. moly-chrome alloy balls is used. The minus 35-mesh product from both classifiers flows to an 8- by 6- by 20-ft. DSF classifier which is close-circuited with two 5- by 9-ft. ball mills using a 2-in. ball charge. The final pulp runs 26 pel' cent plus 150 mesh. The total steel consumption for crushing and grinding is 1.86 lb. per ton of ore milled. Four 35 by 10-ft. DOlT thickeners and one 35- by 15-ft. DOlT balancedtype tray thickener produce pulp underflows, by means of direct-connected -l-iu. Dorrco diaphragm pumps, of 40 pel' cent solids. This thickened

439

TREA'rMENT OF SILVEH ORES

Jewellshaft

l

.

400- tan coarse om bms

t

A/Z.ron feeder

-II--r-

30 cOnJleyor belt

+/-"-

Grizzley I "LSf20Cin{l ---

--T

t.

oJ

Undersize

t

3' TroJ!!or crusher --I

I

t belt 24 "conveyor

r -r I

2-Allis Chalmers lowheadscreens I

t. ~ 24 rnnlle)f0r belt 1/ "f unaersue

Oversize

24 caney"or belt ~4'SYfTlons short head erusher

I:L-

!, 200ton fine oreI bin )

Classifier overflow to flotation 20 cells No. 24 Denver sub-A

2-bonks of 20 cells

3 - 4 disc 6 I American !Yl2e mters

t-'

70 smelter FIG. 96. Flow sheet of the Sunshine mill, Coeur d'Alene district, Idaho.

440


pulp is agitated for 83 hr. in batches in eighteen 15- by "!5-ft. Paohuca tanks and in three 35- by 10-ft. Dorr mechanical agitators. The air pressure in the Pachucas is 35 lb. per sq. in., and about 95 cu. ft. per min. is used. The cyanide is added to the Pachucas to maintain a strength of 4.6 lb. KCN per ton of solution, and lime is held at 0.8 lb. per ton of solution. The cyanide consumption is 2.956 lb. KCN per ton of ore, and the lime consumption is 15.21 lb. of crude lime of 8.15 lb. CaO per ton of ore. Filtration is done in three Merrill center washing slime presses with one hundred 3-in. by 4-ft. by 6-ft. frames, the plates covered with 8-oz. sail canvas, which has a life of 1100 charges or 79 days. The press cycle consists of charging with pulp for 10 min., a barren solution wash of 28 min., a water wash of 34 min. under 55-lb. pressure, and sluicing of the presses for 20 min. with water at 75-lb. pressure on the nozzles. For the washes at the presses 825 tons of barren solution and 1,025 tons of water are used; 2,425 tons of water are used for sluicing the presses. The dissolvedvalues loss in the tailings are 7 cents in silver and 3 cents in gold. The precious metals are precipitated from the solution by means of zinc duct of which 0.4735 lb. per ton of ore milled or 0.03929 lb. per fine oz. of bullion is consumed. The pregnant solution averages about $2.88 per ton, and 2000 tons is precipitated per 24 hr. The effluent carries a trace of the metals. Thirty-five per cent of the silver and 71 per cent of the gold are dissolved in the grinding circuit, 5.5 per cent of the silver and 24 per cent of the gold during agitation and 0.8 pel' cent of the silver and 0.7 per cent of the gold in the filters. In 1948 the Mills-Crowe cyanide recovery process regenerated 174,066 lb. KCN from the barren solution. This enabled the carrying of a high cyanide strength in the agitators, a longer water wash on the Merrill filter presses, and a low mechanical loss in the tailings from the filters. Mochito Mill (Type Vs). This mill, which is located at Mochito, near Lake Yoj oa, treats 100 tons pel' day of a high-grade silver ore carrying about 39 oz. silver pel' ton. It has a relatively high manganese content (4 pel' cent) mostly in the form of pyrolustite, The silver occurs with lead and zinc sulphides. The ore is delivered from the mine to the mill, a distance of about a mile, by means of Diesel trucks. The ore is crushed to a l;!/z-in. product through a No.3 gyratory crusher. Primary grinding is done in a 6 by 5 Allis-Chalmers ball mill in closed circuit with a Denver mineral jig and a 4-ft. by 18-ft. 4-in. Dorr DSFH classifier, and the overflow carries about 10 per cent plus 150 mesh in the pulp. The pulp from the secondary classifier is thickened in a Denver 38- by lO-ft. thickener, the underflow

441


going to eight Massco Fahrenwald flotation cells, five of 'which are used' as roughers, two as cleaners, and one as recleaner. Pine oil, Am'ofloat 31, reagents 404 and 301 are used. The p H is maintained at about 7.6. Flotation tailings are thickened in a 38- by lO-ft. thickener to about 40 per cent solids, the underflow being elevated by Oliver slurry pumps -to two Denver disk filters, 6-ft. diameter rend five leaves each. . The cake from the filters, which carries about 20 per cent moisture, is repulped in barren cyanide solution from the precipitation plant and is then aerated in an 8-ft. Denver agitator, where the lime emulsion is added. Cyanide is then added to the pulp as it flows to agitation. The pulp is agitated 63 hr. in six 12- by 36-ft. Pachuca tanks, the cyanide being maintained at about 6 lb. KCN per ton of solution and the lime at 0.8 lb. The pulp from the Paehucas then flows to four 38- by 10-ft. countercurrent washing thickeners. Two 8- by 10-ft. Oliver filters are now being TABLE

94.

REAGENT CONSUMPTION AND RECOVERY OF VALUES AT J\1CCHITO

Reagent consumption

Per cent of total recovery

Reagent

Lb. per ton milled

Operation

Gold

Cyanide (KCN) ............ Zinc dust .................. Flotation reagents. . . . . . . . . Lime (CaO) ................

7.4 0.66 0.89 12.70

Jigs Flotation Cyanidation Total

41 35.75 86.75

10

Silver

Lead

5 41.2 40.75 86.95

3.5 32.0 35.5

installed for filtration of the pulp from washer 4 in order to lower the mechanical loss in cyanide,' which is high because of the high cyanide strength required during agitation. The precious metals are precipitated from 400 tons of solution daily by the Merrill-Crewe system using zinc dust. Dorado Mill (Type IIa). This mill, located at San Isidro, El Salvador, treats 100 tons pel' day of an oxidized ore carrying 1.61 oz. per ton silver and 0.25 oz. pel' ton gold. The ore is very hard, and about 3.25 lb. of balls is consumed pel' ton in grinding. Due also to the large quantity of wet clay present, it is necessary to wash the are before crushing. The washing is done in cyanide solution by means of a 4- by 19-ft. washing trommel with a I-in, punched-plate screen. The undersize is dewatered in a Stearns-Roger dewatering drag 4 by 26 ft., the overflow of which, containing about 11 pel' cent of the tonnage, is pumped direct to the primary thickener. The trammel oversize is discharged onto a 24-in. picking belt where waste is hand-picked and the are delivered to

442

CYAKIDATION AND CONCENTRA'l'ION OF ORES

a No. 50 Kue-Ken crusher set to crush at 1;~ in. The crusher product plus the dewatered drag sands are delivered to the fine ore bin at the grinding plant by means of a 16-in. by 280-ft., 15-deg. inclined conveyor. Primary grinding is done by means of a 6- by 5-ft. Marcy ball mill in closed circuit with a 3- by 15-ft. Wemco screw classifier. The secondary grinding is carried out in a 4- by 10-ft. ball mill charged with 1.5-in. balls and in closed circuit with a 6-ft. by 21-ft. 4-in. model F DOlT classifier, the overflow being all minus 150-mesh product and flowing directly to the primary thickener, which is 38 ft. 7 in. in diameter by 10 ft. deep. The underflow is maintained at 40 to 45 pel' cent solids and is agitated in three 21-ft. 6-in. by 16-ft. mechanical Wemco agitators with an air pressure of about 12 lb. on the airlift. These are in series and flow into four 38-ft. 7-in. by lO-ft. washing countercurrent thickeners, all underflows being maintained at 40 to 45 pel' cent solids. Six hundred tons of pregnant solution is precipitated daily by means of the Merrill-Crewe system, the precipitates being collected in a bag unit. Owing to the colloidal slimes present in the ore, it has been found necessary to add to the washing solution 0.015 lb. caustic starch and a lime emulsion before washing the ore in order to get flocculation of the slimes. About 10 lb. of lime is consumed in washing the ore. MANGANESE-SILVER ORES

Oxidized silver ores containing the higher oxides of manganese are generally refractory to metallurgical treatment. Manganese fouls mercury if amalgamation of the gold content is attempted. A refractory compound of manganese and silver is formed, probably a manganite, which is insoluble in cyanide solution and other common solvents for silver. Caron Process. The Caron Process (D. S. Patent 1,232,216, Aug. 3, 1917), described by G. H. Clevenger and M. H. Caron in s-a. 226, U.S.B. of 111., is based on the following principle: When oxidized ores containing a refractory compound of manganese and silver are heated in a reducing atmosphere, the higher manganese oxides are reduced to manganous oxides, and if cooled so as to prevent reoxidation, the refractory compound is rendered amenable to cyanidation. Refractory compounds of silver also can be so treated. Manganese-silver ores occur generally in acid-eruptive rocks, chiefly rhyolite and dacite flows of later Tertiary age. Potassium-aluminum silicate is a vein material, and the vein quartz replaces the calcite. The manganese oxide is generally of secondary origin and is formed by atmospheric agencies. For the foregoing reason manganiferous ore from near the surface may be refractory but from depth may be amenable to treat-

'l'REATMEK'J' OF SILVER ORES

443

ment. "Wad," a hydrous manganese manganate, is common in the zone of oxidation. Various treatments of the raw ore have proved unsuitable-concentration (including flotation), magnetic separation, chloridizing, roasting, \Tolatilization, sulphuric acid, and heating with organic matter. The Ag to Mn ratio persists in all sieve sizes from plus 20 to minus 200 mesh. Laboratory tests were made in the United States and in Sumatra, followed by plant-scale runs in the latter country and a 50-ton plant at Pachuca, Mexico. Direct cyanidation of raw ore containing 2 to 10 pel' cent MnO~ gave 50 per cent extraction of the silver, but ore with 25 per cent JVlnO~ gave only 25 per cent extraction. oJ The Caron process, on the other hand, extracted 92 per cent of the gold and 90 per cent of the silver. The pilot plant in MeA-1CO successfully treated ore containing 2.8 to 13 per cent MnO~ and 12 to 20 oz. Ag. The Clevenger kiln (U. S. Patent 1,379,083, May 24, 1921) was fired with producer gas with the following analysis: CO, 15 pel' cent; CI-h 5.5 per cent; H~, 4.6 per cent; CO~, 6 per cent, the remainder being nitrogen. The general conclusions as to the operation and efficacy of the Caron process are 1. Size of ore fed to the rotary kiln may be as coarse as 1 to 2 in. 2. Producer gas of 150 B.t.u. or higher, made from any fuel, may be used. 3. Temperature range is 500 to 700, best at 600°C. 4. Calcine should be discharged into an inert atmosphere or directly into cyanide solution. 5. Unaltered MnO: after calcining should be determined. 6. Alkalinity control is important. 7. Excess air should be used during grinding and agitation. S. Gold in some manganese-silver ores is amenable to direct cyanidation, and its extraction is not increased by reduction. Gold in other ores follows the silver and is more or less refractory and is benefited as much as silver by reduction. 9. Silver extraction of 60 per cent from raw ore may be increased to from SS to 96 per cent. 10. Cyanide consumption is not affected by MnOe. 11. Lime consumption may be 16 lb. per ton. 12. A daily economic minimurn-trentrnent plant for these ores is 200 tons, and the cost could approach $1 per ton, the equivulent of dead roasting and treating a low-sulphur ore.

Caron described the application of the process at Tambang Sawah, Sumatra, Dutch East Indies, a translation appearing in the 111.J. (London) Feb. 19, 1927. The ore, mainly quartz, carried 18 pel' cent manganese dioxide, 30 oz. silver, and 8 dwt. gold. Mter being crushed to 1 in., it was heated 4 hr. ill a rotary kiln. It remained 1;/~ hr. in the reducing section of the kiln in producer gas at 600°C. The JVln02 was reduced to 2 per cent. The ore was then ground and cyanided, yielding 87 per cent of the

4JJ

CYANIDATIO))" AND CONCENTRATION OF ORES

silver and 97 per cent of the gold, as compared with 25 per cent by raw treatment. Chemical consumption was 2.2 lb. cyanide, 5.5 lb. lime, and 1.3 lb. zinc per ton ore. A very complete bibliography covering the treatment of manganesesilver ores is given in Bul. 226, U.S.B. of 111., 1925, by Clevenger and Caron, McClusky Process. Manganese is found in varying percentages in the ores at Fresnillo, Mexico; although, fortunately, the average content is not enough to necessitate special treatment, ores from some parts of the mine contain sufficient manganese to affect seriously the extraction of the silver. To improve the extraction on this relatively small quantity of refractory ore, S. P. McOlusky, formerly metallurgist with the Fresnillo company, developed a modified method of what has become known as the sulphur dioxide process for manganiferous silver ores. This is described by W. E. Crawford in Trans. 112, A.I.M.E., 1934, as follows: Briefly, the method consists of (1) grinding of are in water, (2) subjecting the pulp to the action of sulphur dioxide gas to dissolve the manganese minerals, (3) precipitating the dissolved manganese with a lime emulsion, (4) aerating the pulp, and' finally (5) cyaniding in the usual manner. The ideas involved in the method are (a) that part of the sil vel' is in too close association with the manganese minerals which inhibit the action of cyanide solution on this silver: (b) that, when these minerals are dissolved by S02, the associate.l silver is liberated and thereby becomes accessible to the solvent action of the cyanide solution. Moreover, if the dissolved manganese is then precipitated by lime emulsion and oxidized to the manganic state by aeration, it no longer affects extruction , although it still remains in the pulp. The content of silver in the arc and the gain in extraction by sulphur dioxide treatment, plus the cost of the process, are the criteria by which the applicabili ty of the process to manganiferous silver ores may be judged. At Fresnillo argentite is the predominant silver mineral. It is associated with pyrite and mangaucse minerals. In practice, the ore is ground in a weak, "spent" solution of 0.008 pel' cent KeN so that 30 or 35 pel' cent passes through a 200-mesh sieve and 1 per cent is coarser than 10 mesh. The pulp then flows to a 4-in. Wilfley centrifugal pump, which delivers the pulp to the top of the first and second of the sulphur dioxide treatment towers, which are scaled, airtight chumbers, three in numb e I:. These towers are of wood, 3 by 3 ft. in cross section and 17 ft. high, with wooden baffies lined with whiteiron plates. The purpose of the baffle plates is to disperse the pulp as it falls through the towel', so thut it may come into irrtimute contact wi th the ascending current of SO, gus. The flow of the pulp and SO, gas is countercurrent, the pulp is constantly enriched in acidity and the gas mixture progressively depleted of S02 from unit to unit. The final result is thnt the gas exhausting to the atmosphere contains slightly more than 1 per cent S02 indicating a total absorption of 85 per cent of the avuilable S02. The gus used in this process and in the oyanirle-regeneration plant nt Fresnillo is produced by roasting the pyrite from flotation in seven-hearth Herreshoff furnaces. From the SO, absorption towers the pulp passes through five conditioner tnnks a.rru.nged in series. An emulsion of lime is added to the fourth tank for the purpose of precipitating the dissolved manganous and ferrous compounds, as manganous and ferrous hydrates. The low-pressure air in this and the fifth tank assists in oxidizing

TREATMENT OF SILVER ORES'

445

the manganous and ferrous compoun:ls to manganic and ferric compounds. After passing through the last conditioner tank the pulp is returno.l to the mili for regrinding in a 6- by 14-ft. Traylor ball mill in closed circuit with a DOlT bowl duplex classifier. The overflow of this classifier, which avcrugcs 60 per cent minus 200-mesh material, .i oins the feed of the plant treating the regular sil vel' ore. The gain in extruct.ion necomplisheI by the sulphur dioxide treatment varies 90nsidernbly with different ores, but it appears to be in direct proportion to the amount of manganese dissolved by the gas, approximately 7 grams silver for every 0.1 pel' cent dissolved manganese. An increased recovery by this Lre.rtment, of 25 grams silver per ton, represented a substantial economic advantage when silver was quoted at around 30 cents (United States currency) per ounce. The laboratory pilot test, cnrrierl out daily in conjunction with the plant treatment, often showed as much as 35 grams additional re'covery of silver. Mixing of the S02-treated slimes with the general mill slimes made it difficult to check the actual additional recovery in the plant. An interesting point is noted in connection with tests for the oxygen content of solution in the pulp leaving the final treatment tank of this unit. This solution is entirely devoid of free oxygen; moreover, it required several hours of vigorous agitation with air to satisfy the oxygen-consuming requirement and to render it susceptible to the absorption of free oxygen. In view of this, it is quite possible that a separate cyanide circuit for these treated slimes would be a distinct advantage, especially if it were so designed that several hours of agitation and aeration could be given prior to the addition of cyanide.

CHAPTER

xvn

Costs and Pouter COST OF PLANT CONSTRUCTION

Factors in the cost of building ore-dressing and treatment plants are t}1)e or ore, daily tonnage, accessibility of the mine, and the simplicity or complexity of the process. A mill in 'which amalgamation alone is employed may cost $700 to $1000 pel' ton ore treated pel' day; all-slime cyanide plants, $1200 to $2000 pel' ton; all-flotation plants, $1000 to $1200; cyanideflotation 01' flotation-cyanide plants, at least $1500 pel' ton-day depending upon the size of plant, for in general the unit costs tend to drop as the scale of operation is increased up to about 1000 tons per day. These figures are based on a 50 per cent rise over 1936 costs. On the Rand the prewar cost of a plant consisting of crushers, stamps, tube mills, and sepa rate sand and slime treatment to handle 50,000 tons per month was around £400,000 ($1,900,000), whereas one with crushers, tube mills, and allslime treatment was about £300,000 ($1,400,000) at the old exchange rate. Those in charge of small mines being developed or already equipped to handle up to 100 tons ore daily will find considerable information of value in 1.0.6800, U.S.B. of M., 1934, "Mining and Milling Practices at Small Gold Mines," by E. D. Gardner and C. H. Johnson. Roasting Plants. The prewar cost of roasting plants in Canada, exclusive of the cost of the stack and cyanide equipment but including the building itself, ranged from $2500 to $3500 per ton per day of roaster capacity. Today Edward roaster installations cost from $3000 to $4500 per ton of daily capacity, and it is estimated that the FluoSolids system will cost about $2000 using the same basis of calculation. COST OF OPERATION

Only where a group of mines operates in a single district are costs comparable and then only with reservations. In general, cost systems are fairly uniform, yet in studying costs of a number of plants it is noticeable that in some cases there is a tendency to omit certain operations which are proper charges against ore dressing and treatment. These should covel' the first stage of coarse crushing, whether it be underground or on the surface, as' well as the disposal of the residue, the recovery of bullion, and returns from products sold and must include the cost for labor, power, supplies, repairs, and compensation. In 1936 when data were being compiled for "Cyanidation and Concen446

447

COSTS AND POWER

tration of Gold and Silver Ores," considerable published information was available on milling costs in various parts of the world. At th'e present time, however, it is extremely difficult to obtain reliable figures on the cost of ore treatment owing to the fact that during a period of rising prices and wages the mine managements do not consider current cost data tYiJical of normal operation and are unwilling to release them for publication. Another factor which applies particularly to the United States and Canada .and which tends to make cost-per-ton figures unreliable is the disparity between the rated capacity of many of .,the mills and the actual 500 400 300

\\ , ,

\. ...

" ...

~~

100

80 60 50 40 30

,

~----

/ pre-war r;iqores

20 20 304050 70 100

--

Estimated 1948 figures

-

V

-

200300 500 1,000 Doily capacity; ton: per 24 hours

-

2,000

4,000

FIG. 97. Plot showing relationship between the daily capacity of straight cyanide plants and the over-all cost per ton of are treated. The figures are based on an average of a number of Canadian producers.

tonnage being handled today. This is partly attributed to shortage of underground labor and partly to the fact that during the war period not only was maintenance heavier than normal bu] opportunities for improvements in technique were lacking. ' The operating costs for straight cyanide plants show a greater uniformity than is the case for plants employing combinations of cyanidation and flotation. Figure 97 shows the relationship between the tonnage capacity and total milling cost per ton based on the 1939 figures for a number of typical Canadian plants. Saving in overhead and labor is the principal factor that enters into the decreasing cost per ton for the larger operations. Considerable variation will be found in individual cases depending upon hardness of ore, fineness of grind, hours of treatment required, reagent

448


consumption, and the situation of the property in its bearing on cost of supplies, etc. The total cost of producing an ounce of gold in Canada increased from $22.35 in 1939 to $32.07 in 1945, according to the report of the director of the Ontario Mining Association for 1945. This represents a 43.5 per cent increase. From various other data which are available, however, it appears that milling and treatment costs (mining excluded) have probably not risen on the average over about 30 per cent. The broken line in Fig. 97 indicates estimated present (1948) average cost on the basis of this 30 pel' cent rise. It is probably safe to assume that the milling costs for straight cyanide plants in Canada today (1948) will be found to be somewhere between these two lines. Kerr Addison, for instance, is milling 2800 tons pel' day for a total of 72 cents pel' ton. Hollinger in the 40 weeks ending Oct. 6, 19'18, milled an average of 3627 tons per day at a total cost of 77.29 cents pel' ton, of which 37.90 cents was labor cost. Average figures compiled from a number of plants indicate the following general distribution of costs on a percentage basis, though considerable variations are noted between individual plants.

--

TABLE

95.

AVERAGE DISTRIBUTION OF TREATMENT COSTS

Distribution by Distribution by milling department

Per cent

Crushing and conveying ............ Grinding and classification .......... Cyunidut.ion" ....................... 1 Miscellaneous]. . .. . ................. I Total ...................... .. . . . . . I

15 40 35 10 100

I

labor, power, and supplies

Labor Power Supplies

. . .. . . . .

Per cent

45 20 35 100

'" Where a combination of flotation and cyanidation is used, the combined cost approximates this same percentage. t This includes such items as heating and lighting, sampling, assaying, experimental work, repairs, and various indirect costs, depending upon the system of cost distributions in use. It is partly because widely different methods of charging out such costs have been adopted that considerable divergence in over-all cost distribution is to be found.

DIRECT CYANIDATION

The following typical examples of cost distribution in Canada may be of interest.

4.49

COSTS AND POWER TABLE

96.

COST DISTRIBUTION FOR A lOOO-TON-A-DAY PLANT IN CANADA

Per Cent of Tota! Cost

Sorting. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . Crushing and conveying. .. .. .. .. . .. .. .. .. .. .. .. Grinding and classifying " Thickening and agitation 5.7 Filtering. . , 7.5 Clarifying and precipitating 3.2 Reagents. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. 8.3 Pumping , 5.5 Refining , , J.. 2.0 Total cyaniding. . . . . . . .. . . . . . . . .. .. . .. ... Laboratory and assaying , , 1. 7 Heating , 1.7 Residue disposal. . . . . . . . . . . . . . . . . . . . . . . . . . . . .. 2.3 Mill alterations , 0.1 Total miscellaneous , . . .. .

6.2 15.9 39.7

32.2

5.8 100.0

FLOTATION AND CYANIDATION

A typical example of cost distribution is to be found in the following 1936 figures for a 1000-ton-per-day flotation and cyanidation plant in Ontario, Canada. TABLE

97.

COST DISTRIBUTION FOR A lOOO-TON-A-DAY FLOTATION AND CYANIDATION PLANT IN CANADA

Per Operation -

cent

of total

Operation

--

Per cent of total

Crushing......................... 9.8 Light and heat ..................... 1.7 Ball milling ....................... 19.4 Superintendence ................... 2.2 0.5 Tube milling ....................... 18.5 Repairs .... ................... " Filtering ............... " .... , ... 8.8 Flotation ....................... ... 9.4 Thickening....................... 5.7 Tailings disposal. ................ 2.5 Agitating., ........ ............... 3.7 Reagents ........ . . . . . . . . . . . ....... 8.5 Clarifying and precipitation ...... 2.5 Cleanup, spills, and elevating...... 0.5 Assaying and sampling. " ........ 3.5 Experiments and research ...... ' " . 0.6 TotaL ........... . ...... -.- . ... .. "100.0 Refining. ............ - ............ 2.2 FLOTATION, ROASTING, AND CYANIDATION

Consolidated Beattie gold mines is a good example of a large plant employing flotation, roasting, and the cyanidation of concentrates. Approximately1300. tons per day of arsenical gold is treated for an over-all cost of $1.05 per ton, distributed as shown in Table 98.

450


The roasting cost works out at approximately $1.22 per ton of concentrate, distributed as shown in Table 99. At MacLeod Cockshutt Gold Mines, Ltd., the cost of roasting in 19411942 was 32 cents per ton milled or $1.25 per ton of ore roasted, while at Lake Shore mines for the same year the cost was about 80 cents per ton roasted. TABLE

98.

FLOTATION, ROASTING, AND CYANlDATION COSTS

1948

AT BEATTIE,

Operation

Dollars per ton


Crushing and conveying .. .......... Flotation and drying. .............. Roasting. .. , ....................... Cyanidation ........................

0.236 0.532 0.122 0.157

22.5 50.8 11.7 15.0

Total. ...........................

1.047

100.0

TABLE

99.

ROASTING COST PER TON OF CONCENTRATE

Item

..

Dollars per ton

Per cent distribu tion

Labor .................. Supplies ................ Power .................. General ... ·.·7· • • • • • • • • • • •

0.61 0.23 0.19 0.19

50.0 18.8 15.6 15.6

Total .................

1.22

100.0

LOW-COST OPERATION IN THE UNITED STATES

The 700-ton mill operated by the Standard Cyanide Co. in Nevada between the years 1939 and 1942, when it was closed as a result of government order during the Second World War, succeeded in making a profit from ore carrying as little as 0.06:oz. gold per ton. Cheap, open-pit mining methods were used, and good extractions were obtained when grinding to only 3 mesh. These, among other factors, made for extremely low-cost operation. The 596,482 tons milled yielded $1.86 per ton at a total cost of $1.18 per ton of which $0.52 was milling cost. COSTS ON THE RAND

The distribution of costs at Randfontein Estates Gold Mining Co., which is typical of the older sand-slime type of plant, is shown in Table 100.

451

COSTS AND POWER

In the case of the more modern continuous-treatment plants, East Geduld's total milling cost in 1946 was 36d (60 cents), of which cyaniding accounted for 19.5 cents. The percentage cost distribution at Marievale Consolidated Mines, Ltd., TABLE

100.

OPERATING COSTS-PER CENT DISTRIBUTION AT RANDFONTEIN ESTATES,

1947 Labor'

Power

Operation

Stores White

Native

Sundry

Total

2.81 1.61 3.59 8.35 0.30 1.34 18.00

12.13 16.78 32.97 31.95 1.12 5.05 100.00

Total

--- -oJ

0.74 4.71 Crushing................ 2.50 1.37 3.87 4.67 4.82 Stamping ............... 4.90 0.78 5.68 6.81 17.40 Grinding ................ 5.17 3.18 1.99 14.11 Cyaniding .............. 1.70 7.99 3.80 3.99 0.01 0.30 0.46 0.76 Slime disposal. .......... 10.50 Sand disposal. .......... 0.09 0.48 0.82 2.32 3.14 --- --- --- --41.53 14.37 12.04 26.41 Total. ....... , ........ 14.06

* White labor includes supervision. Native labor includes compound cost, which is about 37 per cent of total native costs. TABLE

101.

PERCENTAGE COST DISTRIBUTION AT MARIEVALE

Labor' Operation

Crushing .................. Milling .................... Cyaniding ................. Totals ...................

Power

1.57 13.53 5.60 20.70

Supplies

10.18 17.35 17.24 44.77

White

Native

6.54 7.24 9.21 22.99

5.85 1.99 2.05 9.89

Misc.

Total

0.71 0.55 0.39 1.65

24.85 40.66 34.49 100.00

* White labor includes operating and engineering, supervision, plant mechanics, foremen, and shift operators. Native labor includes cost of compound which is about 42 per cent total native costs. the newest plant of the Union Corporation, is shown in Table 101 for the year 1947, during which 617,000 tons was milled. POWER

Cyanidation. The power required in cyanide plants varies with type of are, fineness of grind, etc., but in general the range is 20 to 30 kw.-hr. per ton of daily capacity. The power distribution at Preston East Dome mines in Ontario, Canada, is shown in Table 102. The relative distribution of power between the crushing and grinding

452


sections will vary according to the fineness of crushing and the type of plant, but on the average these departmentswill together consume 60 to 70 per cent of the total power. Flotation. The power consumption for straight single-product flotation plants varies, according to A. M. Gaudin,' from 12 to 20 kw.-hr, per TABLE

102.

POWER DISTRIBUTION AT PRESTON EAST DOME

Horsepower

Operation

Sorting (pumping and conveying) Crushing (primary) Crushing (secondary) Screening " Conveying , Magnets Grinding. . Classification Thickening Agitation Filtration Pumping (pulp) '" Pumping (solution) " Low-pressure compressor Vacuum pump Refinery Lights and hot plates TotaL

. . . . . . . . . . . . . . . .

17.5 60 100 10

1

32 10

Per rent distribution

2.4

29.6

J

250 12.5

36.6

4 12

15 33.5

24.5

31

50 30 7.5} 6.9 42 1------'----1----. 717 100.0

103. POWER DISTRIBUTION FOR U.S. FLOTATION PLANTS Crushing, screening, and conveying . .14.6 per cent Grinding and classification . 45.0 Flotation . 28.4 Concentrate disposal . 3.9 Water supply ' . 4.1 Miscellaneous ' . 4.0 100.0 percent

TABLE

ton, depending on the fineness to which the ore is ground. The average percentage power costs for the various. departments of seven United States producers is given in Table 103. POWER CONSUMPTION ON THE RAND

The power consumption at Ranclforitein Estates, which is milling 13,000 tons per day by the older sand-slime process, is shown in Table 104. 1

Flotation, McGraw-Hill, 1932.

453

COSTS AND POWER

Distribution figures for the new 2100-ton-per-clay Marievale plant are shown in Table 105. TABLE

104.

POWER DISTRIBU.TION AT RANDFONTEIN ESTATES,.

Crushing and screening Stamping Primary grinding Secondary grinding Dewatering and pumping water Cyaniding Total. TABLE

105.

. . . . . . .

1946

1.090 kw.vhr. per dry ton milled 6.160 5.888 7.095 0.530 2.790 23.553 kw.vhr. per dry ton milled

.,

POWER DISTRIBUTION AT MARLEVALE

Total crushing plant (including spray water)....... 1.97 kw.vhr. per ton milled Primary grinding. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. 5.48 Secondary grinding. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. 11.60 Dewatering and pumping pulp. . . . . . . . . . . . . . . . . . . .. 0.40 Agitation (including air) 4.00 Filtering. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. 1. 66 Precipitation and solution pumping ~ Total. 26.60 kw.vhr. per ton milled

I

j j j j j j j j j j j j j j j j j j j j j j j J

Appendix

APPENDIX A

Useful Reference Information. This appendix contains miscellaneous information, methods of calculation, factors, and other items to which the millman may refer to refresh his memory. CALCULATION OF RECOVERY IN CONCENTRATE

This method is from E. M. Hamilton's Manual of Cyanidation, 1920: Let C H T R W P L

= assay value of concentrate. = assay value of heads. =

assay value of tails.

= ratio of concentration. = weight of concentrate in per cent.

= percentage recovery.

= percentage loss in tails. R=C-T H-T

W

=

H - T X 100 C-T

P = 100 X C(H - T) H(C - T)

P = C X 100 HXR

L = 100 X T(C - H) "

R(C - T) THE ELEMENTS

Table 106 has been compiled from the Journal of the American Chemical Society, 1933, and from the Handbook of Chemistry and Physics, 1933. (Atomic weights corrected 1950.) . GENERAL CONVERSION FACTORS

One of the most frequent tasks of the engineer is the rapid and accurate conversion of the units of measure of one system into the different but related units of another system. Each engineer remembers the conversion factors of those units that he us~~ most frequently. If, however, he has to convert units other than

457

458


Name

Actinum (rare) ........... Aluminum ............... Antimony ................ Argon (a gas) .......... Arsenic .............. , ... Barium .................. Beryllium ............... Bismuth ................. Boron .................... Bromine (a liquid) .. '" .. Cadmium................ Calcium ................. Carbon .................. Cerium................... Cesium (rare) ............ Chlorine (a gas) .......... Chromium ............... Cobalt ................... Columbium (rare) ........ Copper .................. Dysprosium (rnre )....... Erbium (rare) ............ Europium (rare) ...... " . Fluorine (a gas) .......... Gadolinium (rare) ........ Gallium .................. Germanium .............. Gold ..................... Hafnium (rare) .......... Helium (a gas) ........... Holmium (rare) .......... Hydrogen (a gas) ........ minium (rare) .......... '. Indium (rare) ............ Iodine ................... Iridium .................. Iron ..................... Krypton (a gas) .......... Lanthanum (rare) ........ Lead ..................... Lithium (rare) ........... Lutecium (rare) .......... Magnesium .............. Manganese ............... Masurium ................ Mercury (liquid) ........

106.

Symbol

Ac Al Sb A As Ba Be Bi B Br Cd Ca C Ce Cs CI Cr Co Cb Cu Dy Er Eu F Gd Ga Ge Au Hf He Ho H II In I II' Fe Kr La Pb Li Lu Mg Mn Ma Hg

THE ELEMENTS

Atomic number

Atomic weight

89 13 51 18 33 56 4 83 5 35 48 20 6 58 55 17 24 27 41 29 66 68 63 9 64 31 32 79 72 2 67 1 61 49 53 77 26 36 57 82 3 71 12 25 43 80

227 26.97 121.76 39.944 74.93 137.36 9.02 209.00 10.82 79.916 112.41 40.08 12.00 140.13 132.91 35.497 52.01 58.94 92.91 63.57 162.46 167.64 152.0 19.00 156.9 69.72 72.60 197.2 178.6 4.002 163.5 1.0078 146 114.8 126.92 193.1 55.84 83.7 138.92 207.22 6.940 175.0 24.32 54.93 97.8 200.61

Weight lb. per cu. ft.,

168.5 388.3 86.4 (liquid) 357.7 236.0 114.9 603.7 152.9 194.8 539.9 96.1 140.5 430.7 116.9 94.1 (liquid) 432.0 543.7 524.4 518.1

Melting point, DC.

658.7 630.0 -188 500 850 1280 271 2400 -7.3 320.9 810 3600 700 26 -101.5 1520 1480 1083

298.0 71.2 (liquid) 369.1 340.9 1204.8

-223 29.75 958 1063

9.4 (liquid)

-271

4.4 (liquid)

-259

454.5 308.4 1399.6 438.9 to 493.2 134.8 (liquid) 383.9 687.0 33.3

155 113.5 2350(?) 1530 -157 81O(?) 327 186

108.7 463.2

651 1210

845.6

-39.7

459

APPENDIX A TABLE

Name

Molybdenum ............. Neodymium (rare) ....... Neon (a gas) ............. Nickel. .................. Nitrogen (a gas) ......... Osmium'.................. Oxygen (a gas) ........... Palladium................ Phosphorus (soft) ........ Platinum ................ Polonium (rare) .......... Potassium (soft) ......... Praseodymium (rare) ..... Protoactinium (rare) ..... Radium (rare) ........... Radon (an emanation) ... Rhenium (rare) .......... Rhodium ................. Rubidium (rare) ......... Ruthenium (rare) . .' ...... Samarium (rare) ....... , . Scandium (rare) .......... Selenium.................. Silicon ................... Silver: ................... Sodium (soft) ............ Strontium................ Sulphur .................. Tantalum............. '... Tellurium................ Terbium (rare) ........... Thallium (rare) .......... Thorium ................. Thulium (rare) ...... " ., . Tin ...................... Titanium ................ Tungsten ................ Uranium ................. Vanadium ................ Xenon (a gas) ............ Ytterbium (rare) ......... Yttrium (rare) ........... Zinc ..................... Zirconium ........ '.' ......

106.

THE ELEMENTS

Symbol

Mo Nd' Ne Ni N Os 0 Pd P Pt Po K Pr Pa Ra Rn Re Rh Rb Ru Sa or Sm Sc Be Si Ag Na Sr' S Ta Te Tb Tl Th Tm Sn Ti

W U V Xe Yb Y Zn Zr

(Contimwd)

Atomic number

Atomic weight

42 60 10 28 7 76 8 46 15 78 84 19 59 91 88 86 75 45 37 44 62 21 34 14 47 11 38 16 73 52 65 81 90 69 50 22 74 92 23 54 70 39 30 40

96.0 144.27 20.183 58.69 14.008 191.5 16.000 J 106.7 13.02 195.23 210.0 39.10 140.92 231 226.05 222. 186.31 102.91 85.44 101.7 150.43 45.10 79.96 28.06 107.880 22.997 87.63 32.06 180.88 127.5 159.2 204.39 232.12 169.4 118.70 47.90 184.0 238.07 50.95 131.3 173.04 88.92 65.38 91.22

Weight, lb. per cu. ft.

562.5 434.5 536.9 50.6 (liquid) 1404.6 71.2 (liquid) 759.1 146.1 1334.1 54.3 404.2

Melting point,

'C .

.

2410 840(?) -253(?) 1452 -211 2700 -218 1549 44.2 1755 62.3 940 700

776.6 95.6 752.9 480.7

1950 38 2450(?) 1300

268.4 146.7 650.5 60.6 156.1 124.9 1036.3 375.8

217 1420 960.5 97.5 900(?) 106.8 2900 452

740.4 705.4 455.1 280.9 1161.1 1167.4 372.1 219.7 (liquid) 237.2 448.9 402.0

302 1842 231.9 2000 3400 1850 1720 -140 1490 419.4 1700(?)

460


these, he often has to consult several handbooks before the desired conversion factor is found. With the recognition of a need for a concise table of conversion factors, the following data applicable to metallurgical needs are taken from a compilation by Robert B. Fisher formerly of the Dorr Company for its staff: EXPLANATION 1. Data are-arranged alphabetically. 2. Unless designated otherwise, the British measures of capacity are those used in the United States, ~nd the units of weight and mass are avoirdupois units. 3. The word gallon, used in any conversion factor, designates the United States gallon. To convert into the Imperial gallon, multiply the United States gallon by 0.083267. Likewise, the word ton designates a short ton, 2000 lb. 4. The figures 10-1, 10-', 10- 3 and so on denote 0.1, 0.01, 0.001, respectively. S. The figures 101, 10', 103 and so on denote 10, 100, 1000, respectively. 6. With respect to the properties of water, it freezes. at 32°F. and is at its maximum density at' 39.2°F. In the conversion factors given using the properties of water, calculations are based on water at 39.2°F. in vacuo, weighing 62.427 lb. per cu. ft., or 8.345 lb. per United States gallon. . 7. "Parts per Million," designated as p.p.m., is always by weight and is simply a more convenient method of expressing concentration, either dissolved or undissolved material. As a rule, p.p.m. is used where percentage would be so small as to necessitate several ciphers after the decimal point, as one part per' million is equal to 0.0001 per. cent. 8. As used in the sanitary field, p.p.m. represents the number of pounds of dry solids contained in 1,000,000 lb. of water, including solids. In this field, 1 p.p.m. may be expressed as 8.345 lb. of dry solids to 1,000,000 Vnited States gallons of water. In the metric system, 1 p.p.m. may be expressed as 1 gram of dry solids to 1,000,000 grams of water, or 1 milligram per liter. 9. In arriving at parts per million by means of pounds per million gallons or milligrams per .liter, it may be mentioned that the density of the solution or suspension has been neglected; and if this is appreciably different from unity, the results are slightly in' error.

TABLE 107. CONVERSION FACTORS MuUiply

Acres Acres Acres Acres

By

43,560 4047 1.562 X 10- 3 4840

Acre-feet Acre-feet Acre-feet, Atmospheres

43,560 329,851 1233.49 76.0

Atmospheres Atmospheres

29.92 33.90

To oblain

Square Square Square Square

feet' meters miles yards

Cubic feet Gallons Cubic meters Centimeters of mercury Inches of mercury Feet of water

461

APPENDIX A TABLE

107.

CONVERSION FACTORS

By

Multiply

10,333 14.70

Atmospheres Atmospheres Atmospheres Barrel (oil) (cement) Bags or sacks (cement)

1.058 42 376 94

(Continued) To obtain

Kilograms per square meter Pounds per square inch Tons per square foot Gallons Pounds Pounds

Board-feet British thermal units British thermal units British thermal units

144 sq. in. X 1 in.

British thermal units British thermal units B.t.u. per minute B'-t.u. per minute

107.5 2.928 X 10-4 12.96 0.02356

B.t.u. per minute B.t.u. per minute Centares (centiares) Centigrams

0.01757 17.57

Centiliters Centimeters Centimeters Centimeters

0.01 0.3937 0.01 10

Liters Inches Meters Millimeters

0.2520 777.5 3.927 X 10-4

1

0.01

Cubic inches ., Kilogram-calories Foot-pounds Horsepower-hours Kilogram-meters Kilowatt-hours Foot-pounds per second Horsepower Kilowatts Watts Square meters Grams

Centimeters Centimeters Centimeters Centimeters

of of of of

mercury mercury mercury mercury

0.01316 0.4461 136.0 27.85

Atmospheres Feet of water Kilograms per square meter Pounds per square foot


of mercury per second per second per second

0.1934 1.969 0.03281 0.036

Pounds per square inch Feet per minute Feet per second Kilometers per hour


0.6 persecond 0.02237 per second 3.728 X 10-4 per second 0.03281 per second per second

Cubic Cubic Cubic Cubic

centimeters centimeters centimeters centimeters

3.531 X 10- 5 6.102 X 10-2 10- 6 ' 1.308 X 10- 5

Meters per minute Miles per hour Miles per minute Feet per second per second Cubic Cubic Cubic Cubic

feet inches meters yards

462


Multiply

107.

CONVERSION FACTORS

By

2.642 X 10-;-4 10- 3 2.113 X 10- 3 1.057 X 10- 3


centimeters centimeters centimeters centimeters


feet feet feet feet

2.832 X 104 1728 0.02832 0.03704


feet feet feet feet

7.48052 28.32 59.84 29.92

Cubic feet per minute

472.0

Cubic feet per minute Cubic feet per minute

0.1247 0.4720

Cubic feet per minute

62.43

Cubic feet per second Cubic feet per second

0.646317 448.831

(Continued) To oblain

Gallons Liters Pints (liquid) Quarts (liquid) Cubic Cubic Cubic Cubic

centimeters inches meters yards

Gallons Liters Pints (liquid) Quarts (liquid) Cubic centimeters per second Gallons per second Liters per second Pounds of water per minute Million gallons per day Gallons per minute


inches inches inches inches

16.39 5.787 X 10- 4 1.639 X 10- 5 2.143 X 10- 5



inches inches inches inches

4.329 X 10- 3 1.639 X 10- 2 0.03463 0.01732

Gallons Liters Pints (liquid) Quarts (liquid)


meters meters meters meters

10 5 35.31 61,023 1.308



meters meters meters meters

264.2 103 2113 1057


Cubic yards Cubic yards

7.646 X 105 27

centimeters feet meters yards

centimeters feet inches yards

Cubic centimeters Cubic feet

463

APPENDIX A TABLE

Multiply

107.

(Continued)

CONVERSION FACTORS

By

To obtain

Cubic yards Cubic yards

46,656 0.7646

Cubic inches Cubic meters .


202.0 764.6 1616 807.9


yards yards yards yards

Cubic yards per minute Cubic yards per minute Cubic yards per minute Decigrams

0.45 3.367 12.74 0.1

oJ

Cubic feet per second Gallons per second Liters per second Grams

Deciliters Decimeters Degrees (angle) Degrees (angle)

0.1 0.1 60 0.01745

Liters Meters Minutes Radians

Degrees Degrees Degrees Degrees

3600 0.01745 0.1667 0.002778

Seconds Radians per second Revolutions per minute Revolutions per second

(angle) per second per second per second

10

Dekagrams Dekaliters Dekameters Drams

27.34375

Grams Liters Meters Grains

Drams Drams Fathoms Feet

0.0625 1.771845 6 30.48

Ounces Grams Feet Centimeters

12 0.3048

Inches Meters Yards Atmospheres

10 10

Feet Feet Feet Feet of water

0.02950

Feet of water Feet of water

0.8826 304.8

Feet of water Feet of water

62.43 0.4335

Inches of mercury Kilograms per square meter Pounds per square foot Pounds per square inch

Feet Feet Feet Feet

0.5080 0.01667 0.01829 0.3048

Centimeters per second Feet per second Kilometers per hour Meters per minute

per per per per

minute minute minute minute

%

464


107.

CONVERSION FACTORS

By

Multiply


Feet Feet Feet Feet

per per per per

minute second second second

0.01136 30.48 1.097 0.5921

Miles per hour Centimeters per second Kilometers per hour Knots

Feet Feet Feet Feet

per per per per

second second second second per second

18.29 0.6818 0.01136 30.48

Meters per minute Miles per hour Miles per minute Centimeters per second per second

Feet per second per second

0.3048

Meters per second per second British thermal units Horsepower-hours Kilogram-calories

Foot-pounds Foot-pounds Foot-pounds

1.286 X 10- 3 5.050 X 10- 7 3.241 X 10- 4

Foot-pounds Foot-pounds Foot-pounds per minute

0.1383 3.766 X 10- 7 1.286 X 10- 3

Foot-pounds per minute

0.01667

Foot-pounds per minute Foot-pounds per minute

3.030 X 10- 5 3.241 X 10-,1'

Foot-pounds per minute Foot-pounds per second

2.260 X 10- 5 7.717 X 10- 2

Foot-pounds per second Foot-pounds per second

1.818 X 10- 3 1.945 X 10- 2

Foot-pounds per second Gallons

1.356 X 10- 3 3785

Horsepower Kilogram-calories per minute Kilowatts Cubic centimeters

Gallons Gallons Gallons Gallons

0.1337 231 3.785 X 10-3 4.951 X 10-3


Kilogram-meters Kilowatt-hours British thermal units per minute Foot-pounds per second Horsepower Kilogram-calories per minute Kilowatts British thermal units per minute

feet inches meters yards

Gallons Gallons Gallons Gallons, Imperial

1.20095

Liters Pints (liquid) Quarts (liquid) U. S. gallons

Gallons, U. S. Gallons water

0.83267 8.3453

Imperial gallons Pounds of water

3.785 8 4

.465

APPENDIX A TABLE

107.

Multiply

CONVERSION FACTORS

(Continued)

By

Gallons per minute Gallons per minute Gallons per minute Gallons per minute Gallons water per minute Grains (troy) Grains (troy)

2.228 X 10. 0.06308 8.0208

To obtain. 3

. 8.0208 Area (square feet)

6.0086 1

0.06480

Cubic feet per second Liters per second Cubi c feet per hour Overflow rate (feet per hour) TOIl8 water per 24 hours Grains (avoirdupois) Grams oJ

Grains (troy) Grains (troy) Grams Grams

10- 3 . 103 0,03527 0.03215

Grams Grams Grams Grams Grams Grains per centimeter Grams per cubic centimeter Grams per cubic centimeter Grams Grams Grams Grams

0.01467 2.0833 X 10- 3 980.7 15.43

per per per per

liter liter liter liter

Hectar~s

Hectares Hectograms Hectoliters

2.205 X 10- 3 5.600 X 10- 3 62.43 0.03613 58.417 8.345 0.062427 1000 2.471 1.076 X 105 100 100

Horsepower

100 100 42.44

Horsepower' ,

33,000

Horsepower Horsepower Horsepower

550 1.014 10.70

Horsepower

0.7457

Horsepower Horsepower (boiler)

745.7 33,479

Hectometers I

Hectowatts

Pennyweights (troy) Ounces (troy) Dynes Grains Kilograms Milligrams Ounces Ounces (troy) Pounds Pounds per inch Pounds per cubic foot Pounds per. cubic inch Grains per gallon Pounds per 1000 gallons Pounds per cubic foot Parts per million Acres Square feet Grams Liters Meters Watts British thermal units per minute Foot-pounds per minute Foot-pounds per second Horsepower (metric) Kilogram-calories per minute Kilowatts Watts British thermal units per hour

466


107.

CONVERSION FACTORS

By

Multiply

Horsepower (boiler) Horsepower-hours

9.803 2547

(Continued) To oblain

Kilowatts British thermal units

Horsepower-hours Horsepower-hours Horsepower-hours Horsepower-hours

1.98 X lOG 641.7 2.737 X lOG 0.7457

Inches Inches of mercury Inches of mercury Inches of mercury

2.540 0.03342 1.133 345.3

Centimeters Atmospheres Feet of water Kilograms pel' square meter

Inches Inches Inches Inches

70.73 0.4912 0.002458 0.07355

Pounds pel' square foot Pounds per square inch Atmospheres Inches of mercury

of of of of

mercury mercury water water

Inches of water

25.40

Inches of water Inches of water Inches of water

0.5781 5.202 0.03613

Foot-pounds Kilogram-calories Kilogram-meters Kilowatt-hours

Kilograms per square meter Ounces per square inch Pound per square foot Pound pel' square inch

Kilograms Kilograms Kilograms Kilogrums-

980,665 2.205 1.102 X 10-3 103

Dynes Pounds Tons (short) Grams

Kilogram-calories Kilogram-calories Kilogrum-calories Kilogram-calories

3.968 3086 1.558 X 10-3 1.162 X 10- 3

British thermal unit Foot-pounds Horsepower-hours Kilowatt-hours

Kilogram-oalories per minute Kilogram-calories per minute Kilogram-calories per minute Kilograms per meter Kilograms. per Kilograma per Kilograms per Kilograms per

square meter square meter square meter square meter

Kilograms per square meter Kilograms per square millimeter

51.43 0.09351 0.06972 0.6720

Foot-pounds per second Horsepower Kilowatts Pound per foot

9.678 X lO- G 3.281 X 10-3 2.896 X 10-3 0.2048

Atmospheres Feet of water Inches of mercury Pound pel' square foot

1.422 X 10-3

Pound pel' square inch Kilograms per square meter

lOG

467

APPENDIX A TABLE

107.

CONVERSION FACTORS (Continued)

By

Multiply

To obtain

Kiloliters Kilometers

10' 105

Kilometers Kilometers Kilometers Kilometers

3281 10' 0.6214 1094

Feet Meters Miles Yards

27.78 54.68 0.9113 0.5396

Centimeters pel' second Feet per minute Feet per second Knots Meters per minute Miles per. hour Centimeters per second per second Feet per second per second

Kilometers Kilometers Kilometers Kilometers

per per pel' per

Kilometers Kilometers Kilometers ond Kilometers ond

per hour per hour per hour per sec-

16.67 0.6214 27.78

per hour per sec-

0.9113

Kilometers per hour pel' second Kilowatts

0.2778

hour hour hour hour

56.92

Liters Centimeters

Meters per second per second British thermal units per minute Foot-pounds per minute Foot-pounds pel' second

Kilowatts Kilowatts

4.425 X 104 737.6

Kilowatts Kilowatts

1.341 14.34

Kilowatts Kilowatt-hours

10' 3415

Kilowatt-hours Kilowatt-hours Kilowatt-hours Kilowatt-hours

2.655 X lOG 1.341 860.5 3.671 X 105

Foot-pounds Horsepower-hours Kilogram-calories Kilogram-meters

Liters Liters Liters Liters

10' 0.03531 61.02 10-'

Cubic centimeters Cubic feet Cubic inches Cubic meters

Liters Liters Liters Liters

1.308 X 10-3 0.2642 2.113 1.057

Horsepower Kilogram-calories per minute Watts British thermal units

Cubic yards Gallons Pints (liquid) , Quarts (liquid)

\ 468


107.

CONVERSION FACTORS

By

Multiply

Liters per minute Liters per minute Lumber Width (in.) X thickness (in.)

12


5.886 X 104.403 X 10-3

4

Length (ft.)

Cubic feet per second Gallons per second Board-feet

Meters Meters Meters Meters

100 3.281 39.37 10-3

Centimeters Feet Inches Kilometers

Meters Meters Meters per minute Meters 'per minute

103 1.094 1.667 3.281

Millimeters Yards Centimeters per second Feet per minute


per per per per


per second per second per second per second

0.05468 0.06 0.03728 196.8

minute minute minute second

Meters per second Microns Miles 'Miles

3.281 3.6 0.06 2.237 0.03728 10- 6 1.609 X 105 5280 '.

Feet per second Kilometers per hour Miles per hour Feet per minute Feet per second Kilometers per hour Kilometers per minute Miles per hour Miles per minute Meters Centimeters Feet

Miles Miles Miles per hour Miles per hour

.. 1.609 1760 44.70 88

Kilometers Yards Centimeters per second Feet per minute

Miles Miles Miles Miles

per per per per

hour hour hour hour'

1.467 Ul09 0.8684 26.82

Feet per second Kilometers per hour Knots Meters per minute

Miles Miles Miles Miles

per per per per

minute minute minute minute

Millim's Milligrams

2682 88 1.609 60

Centimeters per second Feet per second Kilometers per minute Miles per hour

103 10-3

Kilograms Grams

APPENDIX A TABLE

107.

CONVERSION FACTORS

By

Multiply

10-3 0.1

Milliliters Millimeters Millimeters Milligrams per liter Million gallons per day Miner's inch

0.03937 1 1.54723 1.5

(Contintted) To obtain

Liters Centimeters '. Inches Parts per million Cubic feet per second Cubic feet per minute

Minutes (angle) Ounces Ounces Ounces

2.909 X 10-4 16 437.5 0.0625

Radians Drams Grains Pounds

Ounces Ounces Ounces Ounces

28.349527 0.!H15 2.790 X 10-5 2.835 X 10- 5

Grams Ounces (troy) Tons (long) Tons. (metric)

Ounces, Ounces, Ounces, Ounces,

480 20 0.08333 31.103481

troy troy troy troy

Ounces, troy Ounces (fluid) Ounces (fluid) Ounces per square inch Pennyweights Pennyweights Pennyweights Pennyweights

24 1. 55517 0.05 4.1667 X 10-3 16 256 7000 0.0005

Pounds Pounds Pounds Pounds Pounds Pounds Pounds Pounds (troy) Pounds Pounds Pounds Pounds

(troy) (troy) (troy) (troy)

1.09714 1.805 0.02957 0.0625


453.5924 1.21528 14.5833 5760 240 12 . 373.24177 . 0.822857

Grains Pennyweights (troy) Pounds (troy) Grams Ounces, avoirdupois Cubic inches Liters Pounds per square inch Grains Grams Ounces (troy) Pounds (troy) Ounces Drams Grains '" Tons (short) Grams Pounds (troy) Ounces (troy Grains

r

Pennyweights (troy) Ounces (troy) Grams Pounds (avoirdupois)

470


TABLE 107.

CONVERSION FACTORS

By

Multiply


Pounds Pounds Pounds Pounds


13.1657 3.6735 X 10-4 4.1143 X 10-4 3.7324 X 10-'1

Ounces (avoirdupois) Tons (long) Tons (short) Tons (metric)

Pounds Pounds Pounds Pounds

of of of of

0.01602 27.68 0.1198 2.670 X 10-4

Cubic feet Cubic inches Gallons Cubic feet per second

water water water water per minute

Pounds per cubic foot

0.01602

Pounds per cubic foot Pounds per cubic foot Pounds per cubic inch

16.02 5.787 X 10-4 27.68

Pounds'per Pounds pel' Pounds per Pounds per

2.768 X 104 1728 1.488 178.6

Kilograms per cubic meter Pounds pel' cubic foot Kilograms per meter Grams per centimeter

Pounds pel' square foot Pounds per square foot

0.01602 4.883

Pounds pel' square foot Pounds per square inch

6.945 X 10-3 0.06804

Feet of water Kilograms pel' square meter Pounds per square inch Atmospheres

Pounds per square inch Pounds per square inch Pounds per square inch

2.307 2.036 703.1

cubic inch cubic inch foot inch

Grams per cubic centimeter Kilograms per cubic meter Pounds per cubic inch Grams per cubic centimeter

Feet of water Inches of mercury Kilograms per square meter Degrees

Quadrants (angle)

90

Quadrants (angle) Quadrants (angle) Quarts (dry) Quarts (liquid)

5400 1.571 67.20 57.75

Minutes Radians Cubic inches Cubic inches

Quintal: Argentine Brazil Castile, Peru

101.28 129.54 101.43

Pounds Pounds Pounds

Chile Mexico Metric Radians

101.41 101.47 220.46 57.30

Pounds Pounds Pounds Degrees

471

APPENDIX A TABLE

107.

CONVERSION FACTORS

By

Multiply


Radians Radians Radians per second Radians per second

3438 0.637 57.30 0.1592

Minutes Quadrants Degrees per second Revolutions per second

Radians per second Radians per second per second

9.549 573.0

Radians per second per second

0.1592

Revolutions per minute Revolutions per minute per minute Revolutions per second per - second Degrees

360

Revolutions Revolutions Revolutions Revolutions per minute Revolutions per minute Revolutions per minute Revolutions minute Revolutions minute Revolutions Revolutions

4 6.283 6

0.1047 0.01667

per minute per

1.745 X 10-3

per minute per

2.778 X 10-4

per second per second

360 6.283 60 6.283

Quadrants Radians Degrees per second Radians per second Revolutions per second Radians per ond Revolutions second Degrees per Radians per

second per secper second per second second

Revolutions per second Revolutions per second per second Revolutions per second per second Seconds (angle)

4.848 X 10- 0

Revolutions per minute Radians per second per second Revolutions per minute per minute Radians

Square centimeters Square centimeters Square centimeters Square ce~timeters

1.076 X 10-3 0.1550 10- 4 100



feet feet feet feet

2.. 296 X 10- 6 929.0 144 0.09290

Acres Square centimeters Square inches Square meters


feet feet inches inches

3.587 X 10-8


3600

3-6 6.452 6.944 X 10-3

feet inches meters millimeters

miles yards centimeters feet

472


107.

Multiply

CONVERSION FACTORS

By

(Continued)· To obtain


inches kilometers kilometers kilometers

645.2 247.1 10.76 X 10° 10°

Square millimeters Acres Square feet Square meters


kilometers kilometers meters meters

0.3861 1.196 X 10° 2.471 X 10- 4 10.76

Square miles Square yards Acres Square feet


meters meters miles miles

3.861 X 10- 7 1.196 640 27.88 X 10°

Square miles Square yards Acres Square feet

Square Square Square S.quare

miles miles millimeters millimeters

2.590 3.098 X 10° 0.01 1.550 X 10- 3



yards yards yards yards

2.066 X 10- 4 9 0.8361 3.228 X 10- 7

Temperature Temperature Temperature Temperature

(Oe.) (Oe.) (OF.) (OF.)

+273 +17.78 +460 -32

1 1.8 1

%

kilometers yards centimeters inches

Acres . Square feet Square meters Square miles Absolute temperature (Oe.) Temperature (OF.) Absolute temperature (OF.) Temperature (Oe.)

Tons Tons Tons Tons

(long) (long) (long) (metric)

1016 2240 1.12000 103

Tons TOllS TOllS Tons

(metric) (ahort ) (short) (short)

2205 2000 32,000 907.18486

Pounds Pounds Ounces Kilograms

2430.56 0.89287 29,166.66 0.90718·

Pounds (troy) Tons (long) Ounces (troy) Tons (metric)

TOllS (short) TOllS (short) TOllS (short) TOllS (short)

Kilograms Pounds Tons (short) Kilograms

I

TOllS of water per 24 hour Tons of water per 24 hour

83.333 0.16643 :

Pounds water per hour Gallons pel' minute

473

AEPENDIX A TABLE

107.

Multiply

Tons of water per 24 hour Watts

Watts Watts Watts Watts

CONVERSION FACTORS

By

1.3349 ·0.05692

44.26 0.7376 1.341 X 10- 3 0.01434.

(Contimwd) To obtain

Cubic feet per hour British thermal" units per minute Foot-pounds per minute Foot-pounds per second Horsepower Kilogram-calories per minute

Watts Watt-hours Watt-hours Watt-hours

10- 3 3.415 2655 1.341 X 10- 3

Watt-hours Watt-hours Watt-hours Yards

0.8605 367.1 10- 3 91.44

Kilogram-calories Kilogram-meters

Yards Yards Yards

3 36 0.9144

Feet Inches Meters

Kilowatts British thermal units Foot-pounds Horsepower-hours

Kilowatt~hours

Centimeters

TABLE

Metal

108.

METALS AND THEIR MINERALS BASED ON INFORMATION FROM DANA'S TEXTBOOK OF MINERALOGY*

Composition

Specific gravity

Formula

Mineral

Hardness

Color and characteristics

~

f!'Antimony ...... Sulphide Cobalt .......... Arsenide Suiph arsenide Copper ......... Metallic Oxide Oxide Carbonate Carbonate Sulphide Sulphide Sulphide Copper ......... Sulphide Antimonide Bulpharsenate Silicate Oxychloride Gold ........... Metallic Telluride Telluride Telluride Graphite ....... Carbon Iron ............ Metallic Oxide Oxide Oxide Carbonate Sulphide Sulphide Sulphide Sulpharsenide Titanate Tungstate Chromate Manganate

Stibnite (antimony glance) Smaltite Cobaltite Native Melaconite (black oxide) Cuprite (red oxide) Malachite Azurite Chalcopyrite Bornite Covellite Chalcocite (copper glance) Tetrahedrite (gray-copper) Enargite Chrysocolla

Sb,S,

4.5 to

CoAs, CoAsS Cu CuO

6.4 6.0 8.8 5.8

CU20 CuCO,Cu(OHh 2 CuCO,.Cu(OH), CuFeS, Cu,FeS, CuS Cu,S

5.8 3.9 3.7 4.1 4.9 4.59 5.5

CuaSb,S,

4.4

2.0

Lead-gray, metallic luster, slightly sectile

to 6.6 to 6.3 to 8.9 to 6.2

5.5 to 6 5.5 2.5 to 3 3

Tin-white, metallic, opaque Silver-white to reddish, cubic, metallic Copper-red, malleable and ductile Black

to 6.1 to 4.0 to 3.8 to 4.3 to 5.4

3.5 3.5 3.5 3.5 3.0

to 5.8

2.5 to 3

Red, brittle Green, brittle Azure-blue Brass-yellow, brittle Copper-red to brown, brittle Indigo-massive Blackish lead-gray, often tarnished, sectile

to 5.1

3

Flint-gray to iron-black

4.6

to to to to

4 4 4 4

o

~ H

ti

~

H

o

!Z ;>

!Z to 4

ti

o

Cu,AsS, CuSiO,

4.4 2 to

+ 2H,O

Atacamite Native Sylvanite Calaverite Petzite Plumbago Native Hematite Limonite Magnetite Siderite Pyrite Marcasite Pyrrhotite

CU2CIH,O, Au (Au.Ag)Te,[Au:Ag (Au.Ag)Te,[Au:Ag (Ag.AuhTe[Au:Ag C Fe Fe20a 2Fe,O" 3H,O FeO, Fe20a FeCO, FeS, FeS, Fe,S, to Fe16S17

Mispickel Ilmenite Wolframite Chromite Franklinite

FeAsS FeTiOa (Fe, Mn)WO, FeC"O, (FeZnMn)O(FeMn),O,

= 1:1] = 6:1] = 1:3]

3.75 15.6 7.9 9.0 8.7 2.0

2.2

3.0 2 to 4

to to to 7 • 4.9 to 3.6 to 5.16 3.8 4.9 to 4.85 to 4.58 to

9.0 2.23 7.8 5.3 4.0

5.1 4.9 4.64

3 2.5 1.5 2.5 2.5 1 4 5.5 5 5.5 3.5 6 6 3.5

5.9 4.5 7.2 4.3 5.1

6.2 5 7.5 4.57 5.22

5.5 5 5 5.5 5.5

to 19.3 to 8.3

to to to to to

to 3.5 to 3 to 2 to to to to to to to to to to

3 2 5 6.5 5.5 6.5 4 6.5 6.5 4.5

to 6 to 6 to 5.5 to 6.5

Grayish-black, brittle Turquoise-blue, translucent, vitreous, rather seetile Bright bottle green, brittle Gold-yellow, malleable and ductile Steel-gray to silver-white to yellow, brittle Pule bronze-yellow I massive Steel-gray to iron-black, brittle Iron-black, dark-steel gray, greasy, flexible Steel-gray to iron black malleable Steel-gray. red Brown, opa que Iron-black, magnetic Gray-brown, reddish, brittle Brass-yellow, brittle Pale bronze-yellow, brittle copper-red, easily Bronze-yellow to tarnished Silver-white, brittle Iron-black, slightly magnetic Brownish-black, brittle Iron and brown black, brittle Iron-black

o

!Z

o

t;I

~

,..,;> H

o

!Z

o>;j o

~ t;I U1

Lead ........... 1 Metallic Carbonate Sulphate Sulphide Mercury........ Metallic Sulphide Molybdenum... Sulphide Oxide NickeL ......... Arsenide Sulphide Silicate Silver........... I Metallic Antimonide Chloride Bromide Sulphide Sulphantimonide Sulpharsenide Sulphantimonide Sulphantimonide Tin............. l Oxide Sulphostannate Tungsten.. . . ... 1 Ferrous Manganous Calcareous Oxide Zinc ............ 1 Oxide Carbonate Sulphide } Zinc blende Silicate Silicate

Native Cerussite Anglesite Galenite Native Cinnabar Molybdenite Molybdite Niccolite } Kupferniekel Millerite Garnierite Native Dyscrasite Cerargyrite (horn silver) Bromyrite Argenite (silver glance) Pyrargvrite (dark ruby silver) Proustite (light ruby silver) Stephanite (brittle silver are) Polybasite Cassi terite Stannite } Tin pyrites Wolframite

Pb PbCO, PbSO, PbS Hg HgS MoS, MoO, NiAs NiS H,(NiMg)SiO, Ag Ag,Sb to Ag,Sb AgCl[Cl 24.7, Ag 75.3]

11.37 6.46 6.3 7.4 13.59 8.0 4.7 4.5

to 6.57 to 7.6 to 8.2 to 4.8

7.3 to 7.67 5.3 2.3 10.1 9.4 5.5

to to to to

5.65 2.8 11.1 9.8

1.5 to 3.5 3 2.75 to 3 2.5 to 2.75 . ............... tc 2.5 2 1 to 1.5 to 2 1 5

to 5.5

to 3.5 3 ................ 2.5 to 3 3.5 to 4 1 to 1.5

Lead-gray White-gray, very brittle White-gray, green, very brittle Lead-gray Tin-white, metallic, brilliant Cochineal-red, sectile Lead-gr~y, foliated, flexible, sectile Straw-yellow, capillary, earthy Pale copper-red, metallic, opaque

Brass tc bronze-yellow, capillary crystals Apple-green, amorphous, soft, friable Massive, filiform, malleable Silver-white, seetile Pearl-gray, resembles wax, highly sectile

................ Bright yellow 2 to 2.5 Lead-gray, sectile

(AgBr) Ag,S[S 12.9, Ag 87.1]

5.8 to 6 7.2 to 7.36

Ag,SbS,

5.7 to 5.8

2.5

Ag,AsS,

5.5 to 5.6

2

to 2.5

Scarlet-vermilion, brittle

Ag,SbS,

6.2 to 6.3

2

to 2.5

Iron-black, brittle

Ag,SbS, SnO,

6.0 to 6.2 6.8 to 7.1

2 6

to 3 tc7

Iron-black Brown, black, infusible

Black, gray-black, brittle

'"d '"d t;;I

Cu,FeSnS, (FeMn)WO,

4.3 to 4.5

4.0

Z

t::J .....

l>1

l>

Steel-gray to iron-black

Dark gray to brown-black, opaque, weakly magnetic ................ ................ Red to hair-brown, nearly black White, yellow, brownish, translucent 5.9 to 6.1 4.5 to 5 ................ ................ Yellow, greenish, earthy 5.4 to 5.7 Deep red, orange-yellow, perfect cleavage 4 to 4.5 4.3 to 4.45 5.0 White, grayish, greenish; brittle 7.2 to 7.5

5

to 5.5

Hiibnerite Scheelite Tungstite Zincite Smithsonite

MnWO, CaWO, WO, ZnO ZnCO,

Sphalerite

ZnS

3.9 to 4.1

3.5 to 4

Yellow, brown, black, perfect cleavage

Calamine Wille mite

H,ZnSip, ZnSiO,

3.4 to 3.5 3.89 to 4.18

4.5 to 5.0 5.5

White, yellowish to brown, brittle White, apple-green, brown

• From Hamilton's Manual of Cyanidation. 1920.

;>

ti I:1l

476


.CONVERSION OF WEIGHTS

Because the conversion of avoirdupois, metric, and troy weights is so frequently done, Table 109, by W. J. Sharwood from Fulton's A 111anual of Fire Assaying, should be of use. TABLE 109. CONVERSION TABLE FOR WEIGHTS (With authentic abbreviations according to Intenationol Critical 'I'ables, Vol. I)

Grains

Weight

Pennyweights

Troy ounces

Avourdupois ounces

Avoirdupois pounds

--I grain (gr.) ......................... 1 0.041666 0.0020833 0.00228571 0.00014285, 1 pennyweight (dwt.). ............. 0.05 24 1 0.0548571 0.00342857 1 1.0071428 10.0685714 1 troy ounce (t. oz.) ................. 480 20 12 13.165714 10.822857 1 troy pound (t. lb.) ... ............. 5,760 240 1 avoirdupois ounce (av, oz.) ....... 437.5 18.2202 0.011458 1 1~·0625 1 avoirdupois pound (av. lb.) ........ 7,000 201. 666 14.58333 16 1 milligram (mg) ......... .... ...... 0.015432 0.000643 0.00000215/ 0.0000352740.00000220'16 1 gram (g) ................ .......... 15.432 0.643 0.03215 0.035274 10.0022046 1 kilogram (kg) ............... ....... 15.432 1643 132.15 35.274 2.2046

Grams --0.064 1.555 31.104 373.248 28.35 453.6 0.001 1 1000

Hundredweight = cwt.

MENSURATION Circle: Area Circumference Diameter Cylinder or prism: Contents Surface Ellipse, area of Parabola, area of Parallelogram, area of Pyramid or cone:* Contents Surface

diameter squared X 0.7854 or radius squared X 3.1416 diameter X 3.1416 circumference X 0.3183 area of end X length or, for a tank, area X depth area of both ends + length X circumference product of two diameters X 0.7854 base X two-thirds altitude base X altitude area of base X one-third altitude circumference of base X half of slant height the base

+ area of

Sphere: Contents diameter cubed X 0.5236 Surface diameter X circumference Wedge with parallel ends, contents of area of base X half altitude * May be used in establishing contents of tailing and ore dumps.

TEMPERATURE-CONVERSION FORMULAS AND VARIOUS TEMPERATURES

1. As the centigrade and Fahrenheit temperature scales have a and 32° as freezing points and 100 and 212 as boiling points, some method of conversion is necessary, and the following formulas are commonly used: 0

477

APPENDIX A

"C. = (OF. - 32) X % of. = DC. X % 32

+

2. Temperatures obtained by various flames and furnaces are !1s follows, according to the Handbook of Chemistry and Physics: TABLE

110.

FLAME AND FURNACE TEMPERATURES

Source

Degrees Centigrade

Industrial furnaces 1700 to Bunsen burner. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Oxycoul gas flame Oxyhydrogen flame , Oxyacetylene flame. . . . . . . . . . . . . . . . . . . . . . . . . . .. Electric arc... . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . ..

1800 1870 2000 2800 3500 3500

3. Temperatures may be judged approximately by the following color scale: TABLE

111.

TEMPERATURE COLOR SCALE

Color

Degrees Centigrade

Incipient red heat })ark-red heat Bright-red heat '. . Yellowish-red heat Incipient white heat VVhite heat

: .. '.

500 to 650 to . . . . . . . . . . . . . . .. 850 to 1050 to 1250 to 1450 to

550 750 950 1150 1350 1550

FORMULA FOR PULP CONSISTENCY

Let p = percentage solids = weight of solids in unit weight of pulp. D = dilution, = water-solid ratio = parts water by weight per part solids. S = specific gravity of dry ore. d = specific gravity of pulp. ,Then p/S = volume of solids in unit weight of pulp. 1 - p = volume of water in unit weight of pulp. 1/d = volume of unit weight of pulp. p 1 and S 1 - p =

+

a:

S

d = p

+ S(l

S _ p)

S -

d=D+1 D+l S

p

=

Sed - 1) d(S - 1)

1

D

+1

pes -

1)

478


S

=

dp 1 - d(l ::- p)

d 1 - D(d - 1)

D=l-p= S-d p Sed - 1)

N ow, let Z

solid factor

=-.!:.-1 P

tons solids per fluid ton (32 cu. ft.) of pulp.

=

= fluid tons of pulp per ton of dry solids. q = percentage solids by volume.

G

62.5d, the weight of 1 fluid ton

Then, as weight of 1 cu. ft. of pulp 32 X 62.5d, and

Z = 32 X 62.5d X p = d 2000 P pS S - peS - 1)

d

-D+1

S DS

+1

Sed - 1) S - 1

By definition, 1

G=-Z G

=

1 pd

=

+

D 1 -d-- ,etc.

By definition,

q

=

2000Z 62.5 X 32S

q

=

d - 1 S - 1

=

Z =

S

p S - peS - 1)

SLIME-DENSITY TABLE

In Metallurgical and Chemical Engineering (now Chemical Engineering) for June, 1912, H. B. Lowden presented the following table and explanatory text for slime-density calculations: Slime-density tables heretofore published have been prepared for use in special cases and are, therefore, not applicable to slimes in which the specfiic gravity of the solids differs from that for which the table was computed. Their value has been chiefly in indicating convenient forms in which the weight and volume relations may be tabulated for use in the control of the cyanide process. The writer, having experienced the need of a more generally applicable table in his work, has prepared one of considerable range with small intervals, which he feels may be useful to others. The table is based on the percentage of solid in the slime, opposite which is given the ratio of solid to liquid. The numbers heading the double columns following are the specific gravities of the dry solid (that of water being taken as unity). The columns headed "S. G." show the specific gravities of the slime, that of water being taken as 1000; that is, the figures show directly the weight of a liter of slime in grams. The columns headed "Vol." show the number of cubic feet of the slime in 1 ton of 2000 lb.

479

APPENDIX A

112.

TABLE

Per cent solids

Specific gravity of pulp and volume of 1 ton in cubic feet, for slimes containing solids of different specific gravities

Ratio of solids to solution

1:19.000 1:15.667 1·: 13.286 1:11.500 1:10.111 1: 9.000 1: 8.091 1: 7.333 1: 6.692 1: 6.144 1: 5.667 1: 5.250 1: 4.882 1: 4.556 1: 4.263 1: 4.000 1: 3.762 1: 3.545 1: 3.348 1: 3.167 1: 3.000 1: 2.846 1: 2.704 1: 2.571 1: 2.448 1: 2.333 1: 2.226 1: 2.125 1: 2.030 1: 1.940 1: 1.857 1: 1. 778 1: 1. 703 1: 1.632 1: 1.564 1: 1.500 1: 1.439 1: 1.381 1: 1.326 1: 1.273 1: 1.222 1: 1.174

2.60

2.50

S.G.

5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46

SLIME-DENSITY RELATIONS

Vol.

B.G.

Vol.

2.70

S.G.

Vol.

2.80

S.G.

Vol.

2.00

S.G.

Vol.

- ---- - -- - - -- - - -- - -

1031 31.03 1037 30.85 1044 30.66 1050 30.46 1057 30.27 1064 30.08 1071 29.88 1078 29.70 1085 29.50 1092 29.31 1099 29.12 1106 28.93 1114 28.74 1121 28.54 1129 28.35 1136 28.17 1144 27.97 1152 27.78 1160 27.58 1168 27.39 1176 27.21 1185. 27.01 1193 26.82 1202 26.62 1211 26.43 1220 26.24 1229 26.05 1238 25.86 1247 25.66 1256 25.47 1266 25.28 1276 25.09 1285 24.90 1295 24.70 1305 24.51 1316 24.32 1326 24.13 1337 23.94 1348 23.74 1359 23.55 1370 23.36 1381 23.17

1032 1036 1045 1052 1059 1065 1073 1080 1087 1094 1102 1109 1117 1125 1133 1140 1148 1157 1165 1173 1182 1190 '1199 1209 1217 1226 1236 1245 1255 1264 1274 1284 1295 1305 1316 1326 1337 1348 1359 1372 1383 1395

31.01 30.82 30.62 30.43 30.23 30.03 29.83 29.64 29.44 29.24 29.05 28.85 28.65 28.45 28.26 28.06 27.87 27.67 27.47 27.27 27.08 26.88 26.68 26.49 26.29 26.10 25.90 25.70 25.50 25.31 25.12 24.91 24.71 24.52 24.32 24.13 23.93 23.73 23.53 23.33 23.14 22.94

1032 1039 1046 1053 1060 1067 1074 1082 1089 1097 1104 1112 1119 1128 1136 1144 1152 1161 1169 1178 1187 1195 1205 1214 1223 1233 1242 1252 1262 1272 1283 1293 1304 1314 1326 1336 1348 1359 1371 1383 1395 1408

31.01 30.79 30.59 30.39 30.19 29.99 29.79 29.59 29.39 29.19 28.99 28.78 28.58 28.38 28.18 27.98 27.77 27.57 27.37 27.17 26.97 26.77 26.56 26.36 26.16 25.95 25.75 25.55 25.35 25.15 24.95 24.75 24.55 24.35 24.14 23.95 23.74 23.55 23.34 23.15 22.94 22.73

1033 1040 1047 1055 1061 1068 1076 1083 1091 1099 1107 1115 1123 1131 1139 1147 1156 1165 1174 1182 1191 1201 1210 1220 1229 1239 1249 1259 1269 1279 1290 1301 1312 1323 1335 1346 1357 1370 1382 1395 1407 1420

30.97 30.76 30.56 30.36 30.15 29.95 29.74 29.53 29.33 29.12 28.91 28.71 28.50 28.30 28.09 27.89 27.68 27.47 27.27 27.06 26.85 26.65 26.44 26.24 26.03 25.83 25.63 25.42 25.21 25.01 24.80 24.60 24.39 24.19 23.98 23.77 23.57 23.36 23.16 22.95 22.74 22.54

1034 1041 1048 1055 1063 1070 1078 1085 1093 1101 1109 1117 1125 1134 1142 1151 1159 1168 1177 1186 1195 1205 1215 1224 1234 1244 1255 1265 1276 1287 1298 1309 1320 1332 1343 1355 1367 1380 1392 1405 1418 1432

30.95 30.74 30.53 30.32 30.11 29.90 29.69 29.48 29.27 29.06 28.85 28.65 28.44 28.23 28.02 27.81 27.60 27.39 27.18 26.97 26.76 26.55 26.34 26.13 25.92 25.71 25.50 25.29 25.08 24.87 24.66 24.45 24.24 24.03 23.82 23.61 23.40 23.19 22..99 22.78 22.57 22.36

480


Per cent

solids

112.

SLIlIIE-DENSITY RELATIONS


Rutio of solids to solution

2.50

8.G.

Vol. --

47 48 49 50 51 52 53 54 55 56 57 58 59 60 61 62 63 64 65 66 67 68 69 70

(Continued)

1: 1: 1: 1: 1: 1: 1: 1: 1: 1. 1: 1: 1: 1: 1: 1: 1: 1: 1: 1: 1: 1: 1: 1:

1.128 1.083 1.041 1.000 0.961 0.923 0.887 0.852 0.819 0.786 0.754 0.724 0.695 0.667 0.639 0.613 0.587 0.563 0.538 0.515 0.493 0.471 0.449 0.429

1393 1404 1416 1429 1441 1453 1466 1479 1493 1506 1520 1534 1548 1563 1577 1592 1608 1623 1639 1656 1672 1689 1706 1724

22.98 22.78 22.59 22.39 22.21 22.02 21.82 21.63 21.44 21.25 21.06 20.86 20.67 20.48 20.29 20.10 19.90 19.71 19.52 19.32 19.14 18.94 18.75 18.56

2.iO

2.60

8.G.

Vol.

8.G.

2.80

Vol.

8.G.

-- -- -- -- -1407 1419 1431 1444 1458 1471 148'1 1498 1512 1526 1540 1555 1572 1585 1601 1617 1633 1650 1667 1684 1701 1719 1738 1757

22.75 22.55 22.35 22.15 21.96 21.76 21.56 21.36 21.17 20.97 20.77 20.58 20.38 20.18 19.98 19.79 19.59 19.40 19.20 19.00 18.80 18.61 18.41 18.21

1420 1433 1446 1460 1473 1487 1501 1515 1530 1545 1560 1574 1591 1607 1623 1641 1657 1675 1692 1711 1730 1749 1768 1786

22.54 22.33 22.13 21.92 21.72 21.52 21.32 21.12 20.92 20.72 20.51 20.31 20.11 19.91 19.71 19.51 19.30 19.10 18.90 18.70 18.50 18.30 18.10 17.90

1433 1446 1460 1473 1488 1502 1516 1532 1547 1563 1579 1595 1611 1628 1645 1662 1681 1698 1718 1738 1757 1776 1797 1818

Vol.

2.00

8.G.

-- --

22.33 22.12 21.92 21.71 21.51 21.30 21.10 20.89 20.69 20.48 20.27 20.07 19.86 19.66 19.45 19.25 19.04 18.84 18.63 18.43 18.22 18.01 1-7.81 17.60

1445 1458 1473 1487 1502 1517 1532 1548 1564 1580 1596 1613 1629 1645 1664 1683 1703 1723 1742 1762 1783 1803 1825 1847

Vol.

-22.15 21.94 21.73 21.52 21.31 21.10 20.89 20.68 20.47 20.26 20.05 19.84 19.63 19.42 19.21 19.00 18.79 18.58 18.37 18.16 17.95 17.74 17.53 17.32

The specific gravities of solids chosen will probably cover the range of slimes ordinarily met with, and the intervals are sufficiently small to admit of interpolation without appreciable error. The last column (4.50) is a hypothetical concentrate and is the specific gravity of a mixture of 80 per cent pyrite and 20 per cent quartz. The average specific gravity of working cyanide solutions is so small as to be negligible. The table is convenient for ascertaining the amount of solid and of solution in slime pulps from the number of cubic feet, determined by rod or float, in the tank; and specific gravity of the slime, determined by taking the weight of a liter or by a specific gravity indicator in the tank. It is useful in calculations for ascertaining the amount of solution to be abstracted or added in thickening and diluting, for correcting the strength of the solutions, for checking tonnage and for other purposes. Assume that in a plant in which the specific gravity of the solid is 2.7, a tank is shown, by the depth of pulp in it, to contain 3530 cu. ft. of pulp, a liter of which weighs 1223 grams. From the table it is found that the specific gravity 1223 corresponds to 26.16 cu ft. per ton and to 29 per cent solid. The weight of pulp; therefore, is 3530 + 26.16 = 135 tons,

481

APPENDIX A TABLE

Per cent

solids

112.

Ratio of solids to solution

17

18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 46

I

3.00

8.G.

1:19.000 1:15.667 1:13.286 1:11.500 1:10.111 1: 9.000 1: 8.091 1: 7.333 1: 6.692 1: 6.144 1: 5.667 1: 5.250 1: 4.882 1: 4.556 1: 4.263 1: 4.000 1: 3.762 1: 3.545 1: 3.348 1: 3.167 1: 3.000 1: 2.846 1: 2.704 1: 2.571 1: 2.448 1: 2.333 1: 2.226 1: 2.125 1: 2.030 1: 1.940 1: 1.857 1: 1. 778 1: 1. 703 1: 1.632 1: 1.564 1: 1.500 1: 1.439 1: 1.381 1: 1.326 1: 1.273 1: 1.222 1: 1.174

(Contin!Led)


Vol.

8.G.

3.20

3.10

Vol.

8.G.

Vol.

3.30

8.G.

Vol.

11.50 f1

8.G.

Vol.

.

-- -- -- -- -- -- - - -- - 1035 30.93 1035 30.92 1036 30.90 1036 30.89 1040 30.76 1042 30.72 1042 30.70 1043 30.68 1043 30.66 1049 30.51 1049 30.51 1049 30.48 1050 30.46 1051 30.43 1058 30.26 1056 30.30 1057 30.27 1058 30,,24 1059 30.21 1067 30.01 1064 30.09 1065 30.05 1066 30.02 1067 29.99 1075 29.76 1071 29.87 1072 29.83 1074 29.80 1075 29.77 1084 29.51 1078 29.65 1080 29.61 1082 29.58 1083 29.54 1093 29.26 1087 29.44 1088 29.40 1090 29.36 1091 29.32 1102 29.01 1095 29.23 1096 29.18 1098 29.14 1099 29.10 1112 28.76 1103 29.01 1105 28.96 1106 28.92 1108 28.88 1122 28.52 1111 28.80 1113 28.74 1115 28.70 1117 28.66 1132 28.27 1119 28.59 1122 28.53 1124 28.48 1125 28.43 1142 28.02 1128 28.37 1130 28.31 1132 28.26 1134 28.21 1152 27·77 1136 28.16 1139 28.10 1141 28.04 1143 27.99 1163 27.52 1145 27.95 1148 27.88 1150 27.82 1153 27.76 1173 27.27 1154 27.73 1157 27.66 1159 27.60 1162 27.54 1184 27.02 1163 27.52 1166 27.44 1169 27.38 1171 27.32 1194 26.77 1172 27.31 1175 27.23 1178 27.16 1181 27.09 1206 26.52 1181 27.09 1184 27.01 1188 26.94 1191 26.87 1218 26.28 1190 26.88 1194 26.79 1198 26.72 1201 26.65 1230 26.03 1200 26.67 1204 26.58 1208 26.50 1211 26.42 1241 25.78 1210 26.45 1214 26.37 1218 26.28 1222 26.20 1253 25.53 1220 26.24 1224 26.15 1228 26.06 1232 25.98 1266 25.28 1230 26.03 1234 25.93 1239 25.84 1242 25.75 1278 25.03 1240 25.81 1244 25.71 1249 25.62 1253 25.53 1291 24.78 1250 25.60 1255 25.50 1260 25.40 1264 25.31 1304 24.53 1261 25.39 1266 25.28 1271 25.18 1275 25.08 1317 24.28 1271 25.17 1277 25.06 1282 24.96 1287 24.86 1331 24.04 1282 24.96 1288 24.85 1293 24.74 1299 24.64 1345 23.79 1293 24.75 1299 24.63 1305 24.52 1311 24.41 1359 23.54 1304 24.53 1310 24.41 1317 24.30 1323 24.19 1374 23.29 1316 24.32 1322 24.19 1329 24.08 1335 23.97 1389 23.04 1328 24.11 1334 23.98 1341 23.86 1347 23.75 1404 22.79 1340 23.89 1346 23.76 1353 23.64 1360 23.52 1420 22.54 1351 23.68 1358 23.55 1366 23.42 1373 23.30 1435 22.29 1363 23.47 1371 23.33 1379 23.20 1387 23.08 1451 22.04 1376 23.26 1384 23.11 1393 22.98 1400 22.85 1468 21.79 1389 23.04 1397 22.89 1406 22.76 1414 22.63 1485 21.55 1402 22.83 1411 22.68 1419 22.54 1428 22.41 1502 21.30 1415 22.61 1425 22.46 1433 22.32 1442 22.18 1519 21.05 1429 22.40 1438 22.24 1447 22.10 1456 21.96 1538 20.80 1443 22.19 1452 22.02 1462 21.88 1471 21.74 1557 20.55

--

5 6 7 8 9 10 11 12 13 14 15 16


482


112.


(Contimted)

Specific gravity of pulp find volume of 1 ton in cubic feet, for slimes containing solids of different specific gravities Per cent solids

Ratio of solids to solution S.G.

47 48 49 50 51 52 53 54 55 56 57 58 59 60 61 62 63 64 65 66· 67 68 69 70 a

1: 1.128 1: 1.083 1: 1.041 1: 1.000 1: 0.961 1: 0.923 1: 0.887 1: 0.852 1: 0.819 1: 0.786 1: 0.754 1: 0.724 1: 0.695 1: 0.667 1: 0.639 1: 0.613 1: 0.587 1: 0.563 1: 0.538 1: 0.515 1: 0.493 1: 0.471 1: 0.449 1: 0.429

3.10

3.00

Vol.

S.G.

3.20

Vol.

S.G.

Vol.

3.30

S.G.

Vol.

4.50"

S.G.

Vol.

---- -- -- -- ---- -- -1457 21.97 1467 21.81 1477 21.66 1487 21.51 1576 20.3o 1471 21.76 1483 21.60 1493 21.44 1503 21.29 1595 20.0 5 1485 21.55 1497 21.38 1508 21.22 1519 21.07 1615 19.81 1500 21.33 1512 21.16 1524 21.00 1535 20.85 1637 19.5 6 1515 21.12 1528 20.94 1540 20.78 1551 20.62 1658 19.3 1 1531 20.91 154'1 20.73 1556 20.56 1568 20.40 1679 19.06 1547 20.69 1560 20.51 1573 20.34 1585 20.18 1700 18.8 1 1563 20.48 1577 20.29 1590 20.12 1603 19.96 1724 18.56 1579 20.27 1594 20.08 1608 19.90 1621 19.73 1748 18.3 1 1596 20.05 1611 19.87 1626 19.68 1640 19.51 1772 18.0 6 1613 19.84 1628 19.65 1645 19.46 1659 19.29 1796 17.8 1 1631 19.63 1646 19.43 1663 19.24 1678 19.06 1822 17.5 6 1649 19.41 1665 19.21 1682 19.02 1697 18.84 1848 17.3 2 1667 19.20 1684 19.00 1702 18.80 1718 18.62 1875 17.07 1686 18.99 1704 18.78 1722 18.58 1739 18.39 1903 16.8 2 1705 18.77 1724 18.56 1742 18.36 1761 18.17 1932 16.57 1724 18.56 1745 18.34 1764 18.14 1783 17.95 1961 16.3 2 1745 18.35 1765 18.12 1786 17.92 1805 17.72 1992 16.0 7 1765 18.13 1786 17.91 1808 17.70 1828 17.50 2023 15.8 2 1786 17.92 1808 17.69 1830 17.48 1852 17.28 2054 15.5 7 1808 17.71 1831 17.47 1853 17.26 1876 17.06 2088 15.3 2 1830 17.49 1854 17.26 1877 17.04 1901 16.83 2123 15.08 1852 17.28 1878 17.0'1 1902 16.82 1927 16.61 2159 14.8 3 1875 17.07 1902 16.83 1926 16.60 1953 16.39 2195 14.5 8

80 per cent pyrite and 20 per cent quartz.

and the weight of solids 135 X 0.29 = 39.15 tons. The weight of solution is, by difference, 95.85tons. If the solution titrates 1.05lb. cyanide per ton and it is desired to bring the strength up to 2.5 lb. per ton, we have 2.5 - 1.05 = 1.45 lb. cyanide to be added per ton. Therefore 95.85 X 1.45 = 139lb. cyanide to be added to the tank. The table is useful in determining the sizes of tanks necessary for any given capacities. Thus, if it is desired to agitate 50 tons of dry slime (specific gravity of solid 2.6) with three parts solution, the table shows this to contain 25 per cent solids and to have a volume of 27.08 cu. ft. per ton; therefore 50 + 0.25 = 200 tons of slime X 27.08 = 5416 cu. ft., the required effective working capacity of the tank, to which an amount must be added to secure the desired height of curb above the charge.

The following logarithmic plot (Fig. 99) is a convenient means for determining particle sizes in inches or millimeters for any screen mesh. (Tyler Standard Scale-.y2.) The meshes are in even figures, and the standard meshes are obtained by interpolation, i.e.. 8, 35, 48, 65, etc.

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Specific 6rOlvi~y of Dry Slime , I [

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Method of Using ChOld If two of three fOlctors be known, the other fClctar is fauna! by plOfCinel CI strOlielhtedge' so tholt it cuts the scotle Olt those reololings. The 'intersection of the strOli(Jht eolge with other sccle gives the reololin£l for the third fOletor

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1000 800 600 500 2

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FIG. 99. Logarithmic plot for determining particle sizes.

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484


TARLE 113. SIEVE SERIES, TYLER STANDARD SCREEN SCALE AND LM.M. SERIES Aperture, inches

Mesh Tyler

I.lILM.

8 10 14 20 28 35 48 65 100 150 200

5 8 12 16 20 30 40 60 100 120 150

Aperture, millimeters

Tyler

LM.M.

Tyler

LM.M.

0.093 0.065 0.046 0.0328 0.0232 0.0164 0.0116 0.0082 0.0058 0.0041 0.0029

0.1 0.062 0.0416 0.0312 0.025 0.0166 0.0125 0.0083 0.005 0.0033 0.0025

2.362 1.651 1.168 0.833 0.589 0.417 0.295 0.208 0.147 0.104 0.074

2.540 1.574 1.056 0.792 0.635 0.421 0.317 0.211 0.127 0.084 0.063

Micron sizes are readily obtained from the millimeter scale by changing the decimal point. (1 millimeter = 1000 microns.) BOOKS ON CYANIDATION ALLAN, A. W.: Mill and Cyanide Hosulbook Charles Griffin & Company Ltd., London, 1918. ALLEN, ROBERT: West Australian 111etalluTgical Practice, Chamber of Mines of Western Australia, 1906. BAIN, H. F.: MOTe Recent Cyanide Practice, Mining and Scientific Press, San Francisco, 1910. BERTELSMANN, WILIIELM: DeT T'echnoloqie der CyanveTbindugen, Munchen and Berlin. 1906. BosQUI, F. L.: Practical Notes on the Cyanide Process, Scientific Publishing Co., New York and London, 1899. CIIARLETON, A. G.: Gold Mining and Milling in Western Australia, Spon and Chamberlin, New York, 1912. CLARK, DONALD: Australian Mining and MetalluTgy, Critchley Parker, Melbourne, Sydney, and Perth, 1904. CLENNELL, J. E.: The ChemistTy of Cyanide Solutions, McGraw-Hill Book Company, Inc., New York, 1904. CLENNELL, J. E.: The Cyanide Handbook, McGraw-Hill Book Company Inc., New York, 1910. EISSLER, MANUEL: The Cyanide Process fOT the 'Extraction of Gold, Crosby Lockwood and Son, London, 1898. FAIIRENWALD, A. W.: The Cyanide Process, McGraw-Hill Book Company, Inc., New York, 1918. . FULTON, C. H.: The Cyanide Process in the Black Hills of South Dakota, Department of Metallurgy, South Dakota School of Mines, 1902. GAZE, W. H.: Practical Handbook of Cyanide Operations, George Robertson and Co., Melbourne, 1898.

APPENDIX A

485

HAMILTON, E. M.: Manual of Cyanidation, McGraw-Hill Book Company, Inc., New York, 1920. Lniernaiional T'extbook: 'The Cyanide Process of Copper and Z.ine Iiefinitu; and S'I1lellin(J, International Textbook Co., Scmnton, Pa., 1911. JAMES, ALFRED: Cyanide Practice, E. and F. N. Span, Ltd" London, 1901. JULIAN, H. F., and EDGAR SMART: Cyanidin(JGold and Silver Ores, Charles Griffin Co., Ltd., London, 1903; revised 1921 by A. W. Allen.

FIG. 100. Bullion pour at Homestake. KUHN, EMIL: Die Chemischen. Vorgange bei del' Cyanlmlglmg von silbererzen, W. Knapp, Halle, 1912. MCCANN, FERDINAND: Cyanide Practice in Mexico (in English), Mining and Scientific Press, San Francisco, 1912. McFARREN, H. W.: Cyanide Practice, McGraw-Hill Book Company, Inc., New York, 1912. MEGRAW, H. A.: Practical Data for the Cyanide Plant, McGraw-Hill Book Company, Inc., New York, 1910. MILLER, A. S.: The Cyanide Process, 2d ed., John Wiley & Sons, Inc., New York, 1906. PARK, JAMES: The Cyanide Process, Champtaloup and Cooper, Auckland, N. Z., and George Robertson and Co., Melbourne, 1896.

486


REID, F. D., J. J. DENNY, and R. H. HUTCHISON: Milling and Metallurgical Practice in Treatment of Silver Ores at Cobalt, Ontario Department of Mines, Canada. 1924. RICKARD, T. A.: Recent Cyanide Practice, Mining and Scientific Press, San Francisco, 1907. ROBINE, R., and lVI. LENGLEN: Cyanide Industry Theoretically and Practically Considered (in French), John Wiley & Sons, Inc., New York, 1906. SCHEIDEL, A.: The Cyanide Process-Its Practical Application and Economical: Results, State Printing Office, Sacramento, 1894. STOKES, RALPH, and others: Rand Metallurgical Practice, Charles Griffin & Co., Ltd., London, 1911-1912. THOMSON, F. A.: Stamp Milling and Cyaniding, McGraw-Hill Book Company, Inc., New York, 1915. . VON BERNEWITZ, M. W.: Cyanide Practice, 1910-1913, Mining and Scientific Press, San Francisco, 1913. WILSON, E. B.: Cyanide Processes, John Wiley & Sons, Inc., New York, 1898.

APPENDIX B

Analytical Methods ANALYSIS OF ORES

The assaying of ores and concentrates for the determination of their gold and silver content is a highly specialized subject which is fully covered in a number of standard textbooks. Among these, the reader is referred to Fire Assaying by O. C. Shepard and W. F. Dietrich and M omual. of Fire Assaying by C. H. Fulton and W. J. Sharwood, both published by the McGraw-Hill Book Company, Inc., New York. ANALYSIS OF CYANIDE SOLUTIONSl

Those concerned with ore testing and plant control are principally interested in the determination of certain component parts of the working solutions that are critical for the proper functioning of the cyanide process. These include especially the free cyanide, protective alkalinity, oxygen content, and reducing power of the cyanide solutions. In special cases it is also desirable to determine the precious-metal content of the solutions and the quantities of various cyanogen compounds that are formed during . ore treatment, such as ferrocyanides, thiocyanates, and base-metal cyanides. While the analysis of gold precipitates and gold bullion is of importance to mill operators, the reader is referred to other sources of information on the subject. Determination of Free Cyanide." Standard silver nitrate solution is made by dissolving 4.33 grams silver nitrate, AgN0 3, in distilled water 1 For a detailed treatment of this subject the reader is referred to J. E. Olennell, The Chemistry of Cyanide Solutions, McGraw-Hill, 1910. 2 It should be pointed out that the presence of certain dissolved impurities in cyanide solution, particuluj-ly soluble sulphides, thiosulphates, and compounds of copper and zinc tends to make this free-cyanide determination unreliable unless special precautions are taken. The interference due to soluble sulphides may be overcome by adding 0.2 to 0.5 grams of litharge or lead carbonate to a 25 ml. solution sample, shaking for a few minutes, and then filtering before titration. In the presence of the cyanide complexes of copper and zinc, varying amounts of the combined cyanide report as free cyanide, depending upon whether or not KI is used and the degree of alkalinity of th~ solution. For copper some authorities recommend using up to 1 gram of K1 for each titration, whereas in the presence of zinc, titration to an opalescent end point without K1 is the more reliable. Where copper and zinc are both present Hamilton advises that titrntions be made both with and without K1 and the lower of the two results taken for solution control.

487

488


and making up to 1 liter. The reaction between silver nitrate and sodium cyanide is represented by the following equation: AgNO:J + 2NaCN = AgNa(CN)2 + NaNO a Thus, 169.9 grams AgNO:J saturates 98 grams NaCN, or 4.33 grams AgNO a saturates 2.5 grams NaCN. One cubic centimeter of the foregoing solution (= 0.00433 gram AgNO a) saturates 0.0025 gram NaCN. Therefore, if 25 cc mill solution be taken for titration, each cubic centimeter of silver nitrate solution required will equal 0.0025 gram NaCN or 0.01 pel' cent NaCN. For example, a 25-cc sample of cyanide solution is titrated, and it is found that 4.8 cc silver nitrate solution is used; the strength of the cyanide solution then is 0.048 per cent NaCN. Sometimes it is more convenient to use a lO-cc sample for titration when strong solutions are used. In such cases the same silver nitrate solution is used, and the burette reading is multiplied by 2.5. The solution to be tested should be clear (filter if necessary). Twentyfive cubic centimeters is put into a clean, transparent, 125-cc Erlenmeyer flask. The solution should not be diluted. The silver nitrate solution is added until the end of the reaction is indicated by the first appearance of a bluish haze, dulling the original brilliancy of the solution. This point is best determined against a black background, and the operator should experiment with various conditions of light until he is able to obtain consistent readings and check himself and others within at least 0.1 cc on the burette. A good setup is to place the burette against a window with a good light (but not in the direct rays of the sun) so that the flask will be about level with the eye against a black background. The addition of a few drops of a 5 per cent neutral solution of KI imparts a yellowish tinge to the precipitate, which makes the exact finishing point more distinct. The effect is due to the precipitation of silver iodide in preference to silver cyanide when the solution contains no more free cyanide. Determination of Total Cyanide. Total cyanide is a term used to indicate, in terms of NaCN (or KCN), all the cyanogen existing in the form of simple cyanides, hydrocyanic acid, and the double cyanide of zinc. Procedure. Measure 25 cc of clear cyanide solution, add 10 cc of caustic soda-potassium iodide solution, and titrate with standard AgNO a solution to the first permanent yellow opalescence. Standard AgNO a solution (see free cyanide above) Caustic Soda (NaOH)-Potassium Iodide (KI) solution: Dissolve 4 grams NaOH and 1 gram KI in 100 cc of water. Determination of Total Cyanide by Distillation.' Twenty-five cubic centimeters of cyanide solution is transferred to a 250-cc Claissen 3 N. Hedley and D. M. Kentro, "Copper Cyanogen Complexes in Cyanidation," Trans. 48, C.M.M., 1945.

APPENDIX B

489

distilling flask leading to two 500-cc Erlenmeyer flasks connected in series. The first of these is placed in an ice bath. Each flask contains' 150 cc of 2 per cent NaOH solution. Fifty cubic centimeters of dilute HCl (1 part 1.16 specific gravity acid to 4 parts water) is added to the Claissen flask through a separatory funnel. The contents of the Claissen flask are then boiled vigorously for about 6 min., after which the stopcock of the separatory funnel is opened to prevent "sucking back" and the flame removed. The contents of the Erlenmeyer flasks are then combined and agitated vigorously for about 1 min. with about 1 gram of litharge to precipitate any small amounts of soluble sulphide present. The solution is then filtered and washed, and the combined filtrate and washings titrated with silver nitrate, using K1 as an indicator. Determination of Protective Alkalinity. Protective alkalinity is usually defined as the "alkaline hydrates and half the monocarbonates." The effect is to protect the cyanide from decomposition by acid constituents of the ore and by carbon dioxide in the atmosphere. Either oxalic acid or a mineral acid may be used as a standard. If oxalic acid be used, the reaction is

Thus, 126 grams H 2C20 4·2H20 saturates 56 grams CaO, or 5.62 grams H 2C20 4·2H20 saturates 2.5 grams CaO. Therefore, if a solution contain':' ing 5.62 grams H 2C2 0 .j · 2H 20 per liter is made up, 1 cc of the solution (= 0.00562 gram oxalic acid) will saturate 0.0025 gram CaO. If 25 cc of mill solution be taken for titration, each cubic centimeter of acid required will equal 0.0025 gram CaO or 0.01 per cent CaO. For example, a 25-cc sample of solution is titrated,and it is found that 3.6 cc standard acid is used. The protective alkalinity of the solution then is equivalent to 0.036 per cent CaO. An oxalic acid solution containing 5.62 grams H 2C2 0 ,· 2H 2 0 per liter has a normality of 0.0892. A mineral acid such as H 2S0 4 or HNO a of the same normality will serve equally well. The acid used should be standardized against a standard alkali, A normal solution is one of which one liter contains a quantity of the substance, expressed in grams, equivalent to one gram hydrogen. When the solution is to be made of a salt that contains water of crystallization, the weight of such water must be taken into consideration. As in the case of oxalic acid.

The total atomic weight, including the water, is 126. Therefore, H 2

= 126, and H = 63 = grams of the salt to be added to 1 liter of water

490


to make a normal solution. A decinormal solution contains one-tenth of this amount, or 6.3 grams per liter. Phenolphthalein (1 gram phenolphthalein in 50 cc alcohol, dilute to 100 cc with water) is used as the indicator, as it gives a value for monocarbonates corresponding with the definition of protective alkalinity already given. The determination (Clennell's method) is made on the same portion of solution used for the free cyanide test as outlined. After the required amount of silver nitrate has been added for the determination of free cyanide, a drop or two of the phenolphthalein indicator is added. If the solution be alkaline, a pink color results. The standard acid is then run in until the pink color disappears, and the burette reading is noted. Determination of Free Lime (CaO). It is important to know the free, or available, CaO in burnt or hydrated limes, especially for the laboratory determination of lime consumption in cyanide tests. The socalled sugar method is a convenient one and is widely used. It is based on the solubility of the CaO present in sugar solution. The carbonates and other oxides are unaffected. Procedure. Add 1.0 gram of the lime, ground to minus 100 mesh, to a 250-cc measuring flask, then add 20 grams of cane sugar and 100 cc water. Shake the flask vigorously for several minutes, then dilute to the 250-cc mark. Let stand at least 2 hr. shaking occasionally, then let settle until the solution is clear. Pipette 25 cc, and titrate, using either sulphuric or oxalic acid as under "Protective Alkalinity," using phenolphthalein as the indicator. The amount of lime (CaO) found by titration multiplied by 10 equals the lime in 1 gram of the sample. Oxygen Content of Cyanide Solutions. Two methods for determining the oxygen content of cyanide solutions are offered as being simple and accurate-that of H. A. White, as described in the J01Lr. C.M. and M .S.s.A., June, 1918, and that of A. J. Weinig and M. W. Bowen, described in Trans. 71, A.I.M.E. 1925. White's method is a colorimetric one, depending on the degree of coloration imparted to a solution of pyrogallic acid in the presence of caustic soda. Weinig and Bowen's method, a modification of that of Schutzenberger, depends on the reducing action of a sodium hydrosulphite solution on a solution of indigo blue (indigotin disulphonate). White Method. This method was developed by H. A. White, consulting metallurgist for the Union Corporation of Johannesburg, and is based upon the color imparted to an alkaline solution on the addition of pyrogallic acid, the "pyro" used in photographic work. ApPARATUS REQUIRED

1 dozen 250-cc glass-stoppered bottles. 1 burette.

APPENDIX B

491

CHEMICALS REQUIRED

Sodium hydrate (NaOH) solution, 2N (80 grams per liter). Pyrogallic acid or pyro (the crystalline salt is preferable to the powder). Brown dye (Diamond brown or caramel). PREPARATION OF STANDARD COLORS

Saturate a quantity of ordinary tap water with oxygen by passing air through it for an hour. Then stand for another hour to remove bubbles of occluded ail'. To one of the 250-cc bottles containing this water add about H 0 gram of pyro and 1 cc 2N N aOH. The pyro crystals must immediately sink below the surface. Then inseft a glass stopper with a twisting motion to exclude any small air bubbles. After the soda solution is added, the bottle should be filled to within % in. of the top, so that the stopper may be inserted at a slight angle. Shake the bottle until all the pyro is dissolved. The water will take on a reddish-brown color corresponding to the oxygen content of oxygen-saturated water at the existing temperature and pressure. The percentage oxygen content of water, saturated under any given set of conditions, may be calculated from the chart of Weinig and Bowen (Fig. 55). This color is then matched with a water solution of Diamond dye or caramel. A small amount of chromate of potash will sometimes assist in obtaining the exact shade. Assume that under the existing conditions the saturation point of water corresponds to 8 milligrams oxygen per liter. Then if standard bottles are made up containing 1 part color' corresponding to saturation and 7 parts water, this lighter color will correspond to 1 milligram oxygen per liter, and equal parts of standard color and water will correspond to 4 milligrams oxygen. In the same way make up a series of eight bottles, colors in which represent oxygen contents of from 1 to 8 milligrams oxygen per liter. TESTING THE SOLUTION

Fill one of the 250-cc bottles with the solution to be tested. Use a rubber tube reaching to the bottom of the bottle, and avoid all agitation. A drop or two of kerosene oil in the bottle will film the solution and still further prevent absorption of oxygen. Fill the bottle nearly full, then add Ho gram pyro and 1 cc 2N NaOH, and stopper instantly, taking care that no air bubble is left under the stopper. Shake well, and compare with standard colors. Some solutions, particularly those resulting from the cyanidation of silver ores, show fading colors and become cloudy on the addition of pyro and soda, according to E. M. Hamilton in E. and M.J., July 17, 1920. In such cases a better comparison is made after the solutions stand a definite

492


time, say 3 to 6 min. Also, in such solutions a better standard color is obtained by making up with a regular plant solution to which is added the usual amount of soda and pyro, Then, after standing 3 to 6 min., this color is matched with the dye or caramel, and the fractional standards prepared as usual. A small amount of solid pigment such as yellow ocher, added to the dye or caramel solution, will match the precipitate which sometimes forms. With such solutions it is better to make up a fresh set of standards whenever oxygen tests are "to be made. Weinig-Bowen Method. The Weinig-Bowen method determines oxygen accurately to tenths of a milligram per liter of solution or 1 part oxygen in 10 million parts of a solution on a 250-cc solution sample, with a proportionately greater degree of accuracy on larger samples. Reasonably clear mill solutions are best sampled by siphoning them through a rubber tube and glass tube into Winchester acid bottles. Pulps should be settled, and the clear liquor siphoned off. A convenient quantity of standard sodium hydrosulphite solution is made up as follows: Fill a 2H-liter acid bottle with distilled water, preferably fresh. Dissolve in it 5 grams caustic soda, and then 'add 5 grams sodium hydrosulphite. Place a layer of kerosene over the solution. Then siphon the liquor into bottle b of Fig. 101. This solution deteriorates rapidly if exposed to the air, so, as shown in the insert a of the sketch, the cork (not rubber) is run in with shellac. The indicator, indigotin disulphonate, is made up as follows: Place in a casserole 7 grams indigotin, and add 30 cc concentrated sulphuric acid. Place over a water bath, and heat to 90°C. for 1H hr. or until all lumps disappear. Then dilute to 2 liters with distilled water. Neutralize the acidity by adding powdered limestone, a little at a time, allowing it to stand a few minutes between additions, until all action has ceased. Filter without washing, place in a corked bottle, and use as required. It is convenient to dilute this solution so that 1 cc of the indicator is equivalent to 0.25 milligram oxygen per liter solution. This will indicate 1 gram per liter when a 250-cc solution sample is taken for titration. This indicator does not deteriorate and may be kept in a well-stoppered bottle. Figure 101 shows the apparatus for this test as follows: two 2H-liter acid bottles a and b, a 250-cc flask c, a 50-cc burette d with side connection, a common burette e, a clamp stand f to hold two burettes, a 400-cc beaker with 250-cc point scratched on it, a glass stirring rod, H6-in. glass or lead tubing and rubber tubing for connections, a pinchcock g for bottom of the rubber connection on the burette that contains standard hydrosulphite solution, and a container for kerosene to be used in the procedure. When setting up the apparatus, the relative positions of the parts shown should be closely observed.

493

APPENDIX B

The bottles are filled as follows: Remove the connection x, and place a cork stopper in the top of the burette so that no solution can overflow. Place a bottle containing 2>~ liters of kerosene so that its bottom is above the top of bottle a, and connect this bottle to the bottom or burette d with a siphon. Open pinchcock g and stopcock i, and allow kerosene to siphon into bottle b until filled. Replace the bottle that contained kerosene by a bottle containing the standard solution of hydrosulphite. This solution should always be covered by a layer of kerosene; siphon the standard solution into bottle b, the kerosene being forced from bottle b over into

Cl

x----------- : :~: c

j f

FIG. 101. Apparatus for determining oxygen in cyanide solutions.

bottle a automatically. As soon as the hydrosulphite solution has reached to within 1 or 2 in. of the top of bottle b, close both the pinchcock g and stopcock j. After flask c has been nearly filled with kerosene, place connection x in top of burette d, and seal with dry shellac dissolved in alcohol. Open stopcock i. keeping pinchcock g closed, and allow the standard solution to pass into burette d until it just enters flask c; then close stopcock i. open pinchcock g, and allow the standard solution to drain completely; its action as a siphon will draw the kerosene over into burette d. The standard solution is now drained off to eliminate any possibility of its being exposed to air and to give it a cover of kerosene in burette d. Close pinchcock g, open stopcock i, and allow burette d to fill to zero mark. The layer of kerosene prevents admission of air during this procedure.

494


Then the apparatus is ready for use. Fill burette e with the indicator, and place a glass covel' over the top to prevent evaporation. Routine titrations are as follows: The solution sample is siphoned over beneath the kerosene into the 400-cc beaker to the 250-cc mark. Alkalinity is neutralized with dilute sulphuric acid, 1 cc or less of indigotin disulphonate is added as indicator, and titration is completed with the hydrosulphite. Then the necessary correction is made for the indicator; and the result is converted into milligrams of oxygen per liter of solution or percentage saturation, as may be desired. The end point in clear solutions is a slight yellow, but with others it may be white to gray, especially if certain salts are contained. The kerosene may be used several times by pouring the contents of the beaker into a large bottle, after titration, then siphoning off the kerosene for reuse after enough has accumulated. General circulating-plant cyanide solutions have 7 to 75 per cent maximum oxygen saturation. A number of precautions must be taken, but these will probably suggest themselves. Determination of Reducing Power.! To 5 to 25 cc of solution, depending upon the amount of reducing agents present, add sufficient water to bring the volume to 200 cc. Then add 25 cc of 25 per cent H 2S0 4 solution and titrate with O.lON potassium permanganate (KMn04) solution to the first faint pink coloration, which remains permanent for 2 min. The result is reported in cubic centimeters of O.lON KMn04 solution per 1000 cc of cyanide solution. The amount of cyanide solution taken for this determination should be so adjusted as to require 5 to 10 cc O.lON KMn04 solution. Larger titrations generally result in fading end points. O.lON KMn04 Solution. Dissolve 3.16 grams of potassium permanganate in water, and dilute to 1000 cc. This solution should be kept in a dark bottle. Determination of Gold and Silver. In the determination of gold and silver in cyanide solutions the degree of accuracy and the speed desired are the governing factors in the choice of methods used and the quanity of solution taken for the determination. 1. Evaporation (Litharge) Method. To an evaporating dish add about 50 grams litharge and 146 to 292 cc cyanide solution. Evaporate to dryness, adding about 10 grams litharge during the evaporation. Scrape out the dried cake, and swab the dish thoroughly with a filter paper moistened with dilute HeI. Flux the cake and paper in a crucible, and cupel the resulting button. ,1 For this and the subsequent determinations described in this section, except where indicated, acknowledgement is made to Analysis of Cyanide Solutions, 2d ed. by The Canadian Industries, Ltd., Montreal, Quebec, 1945.

APPENDIX B

495

This method may be used on either pregnant or barren solutions and is accurate but requires considerable time to complete. 2. Evaporation (Lead Boat) Method. Evaporate sufficient cyanide solution to dryness in a lead-foil boat. When completed, fold the boat inward, hammer into a cube, and cupel. TIlls method is accurate and can be used for both pregnant and barren solutions. 3. Copper Sulphate Method. To 146 to 292 cc cyanide solution add 5 drops of 'saturated potassium ferrocyanide solution, then is cc of precipitating solution, and stir well. Carefully add 20 co H 2S04 , and stir. Filter, wash, and dry the resulting precipitate. Mix the precipitate with the usual assay flux, add the filter paper to the crucible, and cupel the resulting button. This is a generally satisfactory method for both pregnant and barren solutions. Precipitating Solution. The precipitating solution is made as follows: To a saturated solution of copper sulphate add a saturated solution of caustic soda until a light bluish-white heavy precipitate is formed. Add a little more caustic until the color changes to a darker blue, a heavy precipitate remaining. To this add a saturated solution of sodium cyanide until the precipitate is [ust dissolved and a yellow or light brown solution results. 4. Chiddy Method. To 146 to 584 cc (5 to 20 assay tons) cyanide solution add sufficient sodium cyanide, NaCN, to bring the strength to 0.50 per cent NaCN. Add 40 to 50 cc saturated lead acetate, Pb(C 2H,02)2, solution and then 5.0 grams zinc dust, stir well, and heat to boiling. Add 25 cc HCI, and allow to stand on a hot plate until the zinc is dissolved and a sponge forms. Decant the solution, and wash the sponge with water. Squeeze out the excess water, then dry the sponge, wrap in lead foil to bring the weight to about 20 grams, and cupel. NOTE: at least 12 grams of sponge lead should be formed; otherwise the assays are usually low. This method is usually found to be a satisfactory one for routine work, especially on barrens and lower grade solutions. 5. Color Test for Barren Soiuiion». To 1000 cc. of barren solution add 10 cc: saturated sodium cyanide, NaCN, solution. 3 drops saturated lead acetate, Pb(C 2H,02)2, solution. 2 grams zinc dust. Shake well for several minutes. Transfer the sponge to a porcelain casserole, and decant the liquid. Dissolve the sponge in 10 cc aqua regia (3HCI + lHNO a) , and evaporate almost to dryness. Add 5 cc HCI, and evaporate to about 2 cc, Transfer to a test tube, cool, and add care-

496


fully 4 drops of saturated stannous chloride, SnCh, solution. Note the color of the ring formed, and on shake-up note the color of the solution. The following will serve as a guide to the relationship between color and value in cents per ton of the solution. Faint pink (indefinite). Faint blue. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Light purple. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Deep purple. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Deep purple with precipitate. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .. Black Over

2 4 6 8 10 10

It is advisable to make color tests on known barren solutions and establish a color scale to meet the individual mill solution requirements. STANNOUS CHLORIDE (SnCh) SOLUTION. To a 5 per cent HCI solution add stannous chloride until saturated. A few shavings of metallic tin in the solution will keep it in the stannous state. 6. Color Test for Silver. A few drops of a 10 per cent solution of sodium sulphide added to 25 or 50 cc of the plant solution to be tested give a pure white precipitate of zinc sulphide in the absence of silver. The precipitate becomes brownish in the presence of silver, and the depth of color is a close indication of the amount present. The presence of the usual small amounts of lead in the solutions does not affect the result appreciably, according to the Merrill Company. Determination of Ferrocyanide. The most reliable method of determining ferrocyanide in a cyanide solution is to determine the total iron and calculate to ferrocyanide . .Method 1. Volumetric. Procedure. To 200 to 500 cc solution, depending upon the quality of ferro cyanide thought to be present, add 10 cc HCI and 5 cc HNO", and evaporate to about 50 cc. Add 8 cc H 2S0 4, and evaporate to strong fumes and almost dryness. Cool, add 5 cc water and 5 cc H 2S0 4, and again evaporate almost to dryness. Cool, add 50 cc water and 5 cc HCI, and heat to dissolve the soluble salts. Add 3 grams pure zinc (20 mesh). If arsenic is thought to be present, add 3 cc of 4 per cent solution of copper sulphate (CUS04 to 5H20). The copper will deposit on the zinc and the arsenic on the copper as copper arsenide. When the zinc is almost dissolved add 50 cc water and very carefully add 5 cc H 2S0 4. Filter into a 400-cc beaker after all evolution has ceased, wash well with cold water, and dilute the filtrate to 300 cc. Titrate immediately with O.lON potassium permangante, KMn04J solution to a faint permanent pink coloration. O.lON KMn04 Solution. (See "Determination of Reducing Power.")

1 cc O.lON KMn04

=

0.0056 gram iron, Fe

= 0.0369 gram potassium ferrocyanide, K 4Fe(CN)6

497

AP,PENDIX B

1J1ethod 2. Colorimetric. Treat a 100-cc sample of solution as in Method 1, to and including the second evaporation of H 2S04. ' Cool, add 50 cc water, and heat to dissolve the soluble salts. Filter, and wash with hot water. ' To the filtrate, the volume of which should be about 50 cc, add ammonium hydroxide, NH 40H, until alkaline and a slight excess. Heat for a few minutes to coagulate the precipitate, then filter, and wash with hot water. Dissolve the precipitate on the filter with warm 5 per cent H 2S04 solution, and wash with hot water. Add O.ION I\Mn04 solution to a faint pink, avoiding an excess. The pink color will disappear on standi~. Transfer the solution to a Nessler tube, add 10 cc of a 10 per cent solution of potassium thiocyanate, KCNS, and dilute to lOQ-cc mark with water. To the second Nessler tube add 75 cc of 5 per cent.I:J;2~04 solution and 10 cc of the KCNS solution. Then add from a burette drop by drop the standard iron solution until the colors match, then,"add sufficient water to bring the solution to the even IOO-cc mark. " .. STANDARD IRON SOLUTION. To 200 cc of 5 pe{cent solution of H 2S04 add 7.021 grams of ferrous ammonium sulphate' (FeS04 (NH4)2S04 to 6H 20}. Add O.lON KMn04 to a faint pink coloration, then add 40 cc H 2S04, and dilute to 1000 cc. ('

, '

0.001 gram Fe

Determination of Thiocyanate. Method 1. Colorimetric. To a 100cc Nessler tube add 50 cc water and 5 cc (more if necessary) of the cyanide solution to be tested, then add 2 cc HCI and 10 cc of 5 per cent solution of ferric chloride, FeCb. Mix, and dilute to the 100-cc mark with water. If Prussian blue should form on the addition of FeCL, the solution should be filtered and the precipitate washed with water until all the red color is washed from the paper. Return the filtrate to the tube, and dilute to the IOO-cc mark. To a second Nessler tube add 50 cc water, 2 cc HCl and 10 cc 5 per cent FeCI, solution. Then dilute with water to almost the IOO-cc mark. From a burette run in O.OIN potassium thiocyanate, KCNS, solution a few drops at a time until the color in tube 2 matches that in tube 1. When matched, read the burette, and calculate the amount of KCNS added to effect the match. O.01N KCNS SOLUTION. Dissolve 9.72 grams chemically. pure KCNS in water, and dilute to 1000 cc. This is O.lON solution. Mix well, then take 100 cc of this solution, and dilute to 1000 cc. This is .01ON solution. 1 cc = 0.000972 gram KCNS

498


The burette reading X 0.000972 = weight in grams of KONS in the original sample. FIVE PER OENT FeOL SOLUTION. To about 100 cc water add 50 grams FeOl, and 25 cc HOI. Warm to dissolve the FeOL. 0001, and dilute to 1000 cc. This method is sufficiently accurate for most purposes and may be done in a few minutes. JJ;[ethod 2. Permanqamaie ~Method. To 100 cc of cyanide solution add 10 cc H 2S04 and 10 cc 20 per cent solution of ferric sulphate, Fe2(S04)a. Stir well, and filter into a 100-cc graduate cylinder until the 60-cc mark is reached. The filtrate represents 50 cc of the original solution. Wash into a 250-cc beaker, dilute to about 100 cc, and titrate with 0.10N potassium permanganate, KMn04, solution until the red color is dispelled. 1 cc O.lON KMnO = 0.00162 gram potassium thiocyanate, KONS On 60 cc aliquot = 0.0324 gram per 1000 cc or 0.0648 lb. per ton of solution

O.lON KMn04 SOLUTION. (See under "Determination of Reducing Power.") Silver interferes, but it is unlikely to be found in sufficient quantities in the cyanidation of gold ores to cause an appreciable error. Determination of Copper. Method 1. Short Iodide. To 200 to 500 cc of solution add 10 cc HOI, 5 cc HN0 3. Evaporate to about 50 cc, then cool, and add 8 cc H 2S04. Evaporate almost to dryness. 0001, add 5 cc water and 5 cc H 2S04, and again evaporate almost to dryness. 0001, add 50 cc water, and heat to boiling. Add 5 cc concentrated bromine water, and boil to expel the bromine. 0001, add ammonium hydroxide, NH40H, drop by drop, until the ferric hydroxide precipitate just remains after stirring. Do not add an excess of NH 40H. Add' 5 cc acetic acid, H(02H302), and heat, stirring until the iron precipitate dissolves and the solution turns to a clear wine color. Add 1 to 2 grams sodium fluoride, NaF, the solution turning blue. Stir well, cool, and add 4 cc of 50 per cent solution of potassium iodide, K1, solution turning brown, then add a few cubic centimeters of starch solution, and titrate carefully, but quickly, with the standard sodium thiosulphate, N a 2S203: 5H 20, solution. The end point is usually sharp, the color changing from blue to a creamy white. STANDARD SODIUM THIOSULPHATE SOLUTION. Dissolve 19.5 grams sodium thiosulphate crystals, Na 2S20a to 5H 20, and 2 grams NaOH in a convenient amount of water, and then dilute to 1000 cc. Solution should be kept in a dark-colored bottle. 1 cc N a2S20a to 5H 20 solution = 0.005 gram copper, Cu On a 200-cc sample 1 cc = 0.025 gram per liter or 0.050 lb. per ton of solution

APPENDIX

B

499

STARCH SOLUTION. To 1.0 gram of soluble starch add sufficient water to make a paste. Then add 100 cc warm water and 0.1 gram NaOH, and bring to a boil. This solution should keep for several weeks. It/Iethod 2. Colorimetric. To 100 cc of solution add acids, and treat as in Method 1 to and including the second addition of H 2S0 4. Evaporate, andfume strongly until only 2 to 3 cc remain. Cool, add 30 cc water, and heat to dissolve the soluble salts. Filter, and wash twice with hot water. Add 50 per cent ammonium hydroxide, NH 40H, solution until just neutralized, then .add 10 cc more. Warm (do not boil) until the fine precipitate just coagulates, and filter through a fine filter paper. Wash three times with hot water, then transfer to 100-cc Nessler tube, and dilute with water to the 100-cc mark. To a second Nessler tube add 80 cc water, 10 cc NH40H. Add the standard copper sulphate solution 1 cc at a time until the colors match. STANDARD COPPER SULPHATE SOLUTION. Dissolve 0.393 gram copper sulphate, CUS04' 5H 20, in water, and dilute to 1000 cc. 1 cc = 0.0001 gram copper, Cu On 100-cc sample 1 cc = 0.001 gram Cu per 1000 cc or 0.002 lb. per ton of solution This is a satisfactory method for small amounts of copper up to 0.010 gram per liter. If the copper content exceeds this figure, the method may be used on 50- or 25-cc samples. Determination of Zinc. Zinc usually occurs in cyanide solutions as the double cyanide, but under certain conditions, e.g., in dilute solutions, a portion of the zinc may be present as zinc cyanide. It is possible that some may also exist as an alkaline zincate. Procedure. To 500 cc of solution add 10 cc HCl, 10 cc HNO a, and 8 cc H 2S0 4. Evaporate on a hot plate until copious fumes of S03 are evolved. Take down over a flame until 2 to 3 cc remain. Take up with 10 cc water and 5 cc H 2S0 4, and again evaporate to strong fumes of S03 and dryness. Care should betaken at this stage to avoid spitting. Cool, add 100 cc water and 8 cc H 2S0 4, and boil until all soluble salts are in solution. The volume of solution should be maintained at about 100 cc by adding water as required. Add about 6 grams of aluminum either as 30 mesh or sheet aluminum, bent up at the corners, and boil for 10 min. The aluminum precipitates copper, lead, nickel, cobalt, arsenic, antimony, and tin. Filter, and wash with hot water several times. Some extra aluminum should be present in the filter paper. Evaporate the filtrate to about 40 cc, and add 10 cc HN0 3 and 1 gram potassium chlorate, KCl0 3. Evaporate to dryness. Do not bake the residue. Add 50 cc water and H gram N aOH. Break

500


up the cake on the bottom of the beaker, and then add 7 grams ammonium carbonate, (NH 4)2CO", and heat nearly to boiling for several minutes. Let the precipitate settle, filter, and wash several times with hot 5 per cent (NH 4)2CO" solution. Make the filtrate acid with HCl, and add 3 cc in excess, then dilute to 200 cc. Heat to 70°C., and titrate very slowly with potassium ferrocyanide, K~e(CN)o, solution, using uranium acetate as an outside indicator. The end point is reached when a drop of the solution, placed on a white tile, shows a brown tinge when touched with a drop of the uranyl acetate after standing about a minute. The solution should be kept at about 70°C. during titration and be constantly stirred. STANDARD POTASSIUM FERROCYANIDE, K 4Fe(CN)o, SOLUTION. Dissolve 21.6 grams chemically pure crystals, K 4Fe(CN)o to 3H20 , in water and dilute to 1000 cc. 1 cc K 4Fe(CN)o solution = 0.005 gram zinc, Zn URANIUM (URANYL) ACETATE INDICATOR. Three grams uranium acetate and >'2 cc acetic acid in 100 cc water. In order to save time in titrating, the zinc solution may be divided into two nearly equal parts. Titrate one of these parts to an approximate end point, then add the remainder of the solution, and finish the titration. For the determination of other elements, such as antimony, arsenic, cobalt, nickel, lead, soluble sulphates and sulphides in cyanide solutions the reader is referred again to "Analysis of Cyanide Solutions" by the Canadian Industries Ltd. Qualitative Test for Traces of Cyanide.' To 500 to 1000 cc of the solution to be tested add 1 to 2 cc ammonium sulphide, (NH 4)2S, and evaporate just to dryness. The final stages of evaporation should be done slowly. Cool, add 10 cc water, stir well, let settle, and filter. To the filtrate add 2 drops of saturated ferric chloride, FeC13, solution. A red coloration indicates the presence of cyanide. Analysis of Gold Precipitate. The predominating elements in gold precipitates are gold, silver, lead, and zinc. The principal secondary elements may include all or any of the following: arsenic, antimony, copper, iron, nickel, sulphur, insoluble lime; and tellurium. Methods of determination are described on pages 19-23 of the Canadian Industries, Ltd., publication, Analysis of Cyanide Solutions, 2d ed., 1945. 6 There is a tendency for copper, cobalt, and mercury to interfere in this method. For a detailed discussion of various quantitative methods for the determination of traces of cyanide reference should be made to "Cyanide in Mill Effluents" by R. E. Rickard, Mining Mag., London, July, 1938.

s

The Westport Mill-research and testing laboratories of the Dorr Company at Westport, Connecticut.

I

j j j j j j j j j j j j j j j j j j j j j j j j j j j j j j j J

Index A

Achotla chloridizing mill, 272-273 Acid brine, leaching with, 247 Acid treatment of filter cloth, 110 Addition agents, 38-39 Adsorbed .gold and silver, recovery of, 263-265 Aeration, of calcine, 171-172 of mill solutions, 217-218 of sand charge, 91 Aeration tests, 29 Aero-brand cyanide, 210 Agitation, continuous vs. batch, 106-107 theory of, 102-103 Agitators, Devereux, 103,217 Dorr, 103-105,217 Noranda, 103, 105-106 Pachuca, 102, 104, 221 Turbo and Wallace, 103-104,217 Akins, classifier, 75, 80, 90 Alaska, gold production of, 324 Alkalinity, control of, 213, 239-244 Altaite, 281 Amalgam, gold, 177 ' sodium, 28 Amalgamation, at Argonaut, 178-179 barrel, 181-182 use of chemicals for, 183 and chloridising, 271-272 direct, 128, 177-181 at Homestake, 179-181 at Pickle Crowe, 183-184 plate, 138-139, 178-179 at Porcupine United, 284-286 principles of, 177 Amalgamation tests, 27-28 Ammonia-cyanide process, 268 Analysis, chemical, of concentrate at gold mines of Kalgurli , 412 of cyanide solutions, 487-500 of gold bullion, 487 of gold precipitate, 500 of Me Intyre mill products, 312 of mill feed, 410 of Pachuca ore, 432 at Randfontein, 361

Analysis, of Sunshine concentrate,. 438 Antimony ores, roasting of, 250 treatment of, 249-250 Aqua Fria, 342 Ariston, 386-387 Arsenic ores, treatment of, 248-249, 267 Arsenopyrite, 242 flotation df, 146 washing of, 158, 162-168 Arsenopyrite ores, treatment of, 248-267 Ashanti , 386, 387 precipitate treatment at, 202 Assaying (see Fire assaying) AtokBig Wedge, 421-423, 425 Atomic weights of the elements, 458-459 Australia, 396-415 Automatic recording and controls, 234237

B Banket ore, 219 Barren solution (see Solution) Beattie, costs at, 449-450 roasting at, 162-164 Belgian Congo, 396 Bendelari jig, 134 Berdan pan, 182 Bibiani, 388-392 precipitate treatment at, 201-202 Bidi, Saruwak, 92-95 Big Bell, 398-399 Bird centrifuge, 85-86 Blanket strakes (see Corduroy) Books on cyanidation, 484-486 Brazil, 347 British Guiana, 352 British West Africa (see Gold Coast) Bromocyanide process, 266-267 for arsenic, tellurides, 267 Bromocyanide tests, 39 Buffer salts, use of, 244 Bullion refining, at Hollinger, 204-205 at Homestake, 205-206 Burma, 426 Bustiek Mines, Ltd., 385 Butters-Mein distributors, 88 503

504


C Calcine treatment, 171-176, 269 gold losses in, 174-176 at Lake Shore, 171-173 at Rietfontein, 173-174 Caldecott cones, 79-80 Calera, 346-347 California, gold mining in, 331-333 Cam and Motor Gold Mining Co., Ltd., 385 Canada, gold production in, 283-284 milling in, costs of, 447-450 method of, 281-283 ore deposits in, 281 Canvas, 134 Carbon cyanidation, 261-265 at Getchell, 262-203, 265 at Harquahala, 202-:;:04. revolving screens for, 262-263 Carbonaceous matter, 149-150 in California ores, 333 (See also Graphite ores) Cariboo Gold Quartz Mining Co., Ltd., 294-296 Caron Processes, .442-444 Caustic soda, use of, in ion-exchange, 266 CCD (counter current decantation), principles of, 122 vs. stage filtration, 294-296 uses and calculation of, 122-126 Centrifuge, Bird, 85-86 Chaffers retreatment plant, 403-404 Chalcopyrite, 240-241, 244, 274 Chalmersite, 348 Chili, 351-352 China and Manchuria, 427 Chloridizing and amalgamation, 271-272 Chloridizing roasting and leaching, 272273 Chlorination, 270-271 of waste cyanide solution, 275-276 Chlorine, as a metallurgical agent, and as a solvent for gold, 269 Chromium minerals, effect of, 247-248 Clarification of solution, 184-186 Clark-Todd amalgamators, 79, 179-181 Classifier-grinding circuits, 77 Classifiers, bowl, 76-77, 87 centrifugal, 85-86 cone, 79-80

Classifiers, hydraulic, 83 reciprocating rake, 81 screw, 80 Classification, control of, 236 definition of, 74 for leaching, 87-88, 90 means of, 78-79 of mill flow sheets, 279-281 uses of, 75-76 Cochenour Willans Gold Mines, Ltd., 318-320 roasting at, 166-168 Coco-matting, 134 Columbia, 342-343 Compression belts and rolls, 120 Comstock lode, 333-334 Concentrate, smelting, 155-157 treatment of, 153-157 at Idaho-Maryland, 336, 338-339 in Ontario, 154-155 Concentration, 128-157 recovery calculation of, 457 Conditioning ore pulp, 220-222 for cyanidation, 221 for flotation, 221-222 Cone, classifying, 79-80 gold trapping, 130 Conversion factors, 457-476 Copper, determinations of, 498-499 removal of, by leaching, 268-269 Copper complexes, 244-245, 260-261, 268 Copper minerals, 474-475 effect of, in cyanidation, 22G-244 Corduroy, at Dome Mine, 137-138 as a gold saver, 134-136 method of applying, 137 at Modderfontein East, 139 on the Rand, 138-139 Corduroy tests, 28 Costs, of plant construction, 446 of plant operation, 446-451 on Witwatersrand, 357, 450-451 Cripple Creek, 327 Croesus Proprietary Treatment Co., Ltd., 404-405 Crushers, fe-eding, 49 gyratory type, 47, 48-51 jaw type, 46, 48-50 roll, 46, 50 Crushing, at Hollinger, 51-52

505

INDEX

Crushing, at Me Intyre, 52-53 surface, 50 underground, 49-50 Cyanide, consumption of, 211, 215, 235 determination of, 36-37 determination of, 487-489, 500 history and manufacture of, 209 regeneration, processes for, 253, 259, 261 in Australia, 260-261 at Flin Flon, 256-258 at Pachuca, 255-256 Cyanide plants, cost of construction, 446 cost of operation, 446-451 Cyanide solution, aeration of, 217-218 alkalinity of, 213, 239-244 analysis of, 235 (See also Appendix B) impurities in, 218-220, 233 oxygen in, 215-216 reducing agents in, 216-217 strength of, 210, 243 temperature control of, 243 (See also Solution) toxicity of, 275 Cyanide solution waste, treatment of, 275-276 Cyanide tests, by agitation, 35-40 by percolation, 40-42 Cyclone (see DSM cyclone) Cyanicides, control of, at Homestake, 242-243 at Lake Shore, 245 at Morro Velho, 243-244 at N oranda, 240-245 at Salsigne, 241-242 at Sub-Nigel, 240-241

D De-aeration of solution, 186 Deister concentrator, 140 Density of the elements, 458-459 Denver flotation machine, 144-145 Denver mineral jig, 134 Desliming, 84 Determination, of copper, 498-499 of cyanide, free, 487-488 total, 488-489 traces of, 500

Determination, of ferrocyanide, 496-497 of free lime, 490 of gold and silver in solution; 494-496 of oxygen in cyanide solutions, 490-4940 of protective alkalinity, 489-490 of reducing power, 494 of thiocyanate, 497-498 of zinc, 499-500 Dewaterers, mechanical, 80 Diaphragm pump, 100-101 Dome mines,.288-291 concentrate treatment at, 154 conditioning at, 221 desliming at, 85 Dorado mill, 411-442 Dorr agitator, 103-105,217 for carbon cyanidation, 263 Dorr classifier, 75-76, 79-82, 90 Dorr thickener, 96-100 Dorrco filter, 116-118 Dorrco Pan-American jig, 133-134 Dorrco sizer, 83-84, 142,236 DSM (Dutch State Mines) cyclone, 86 Dust lubrication, 161

E East-Geduld, 370-373 costs at, 451 Ecuador, 344-345 Edquist process, 200 Edwards furnace, 159-162, 165 installation cost for, 446 EI Callao, 351 Elements, table of the, 458-459 Examination of are, 26-27

F

L

Fagergren flotation machine, 144, 147 Feeders, apron, Ross chain, 49 automatic, 235 Ferroeyanide, determination of, 496-497 formation of, 239-240, 243 Ferrous sulphate, 240 Fiji Islands, 415-421 Filter cloth, 108-109 acid treatment of, 118-119 Filter tests, 44-45

I

I

506


Filters, precipitate, bag type, 190 pressure type, 190-191 vacuum type, 192-193 pulp, types of, 110-119 Filtration, definition of, 107-108 principles of, 108 stage, 110 . Filtration media, 108-109 Fire assaying, 487 Flappers, use of, 120 Flocculating reagents, 44 Flotation, automatic control of, 236-237 conditioning for, 221 of cyanide residues, 152-153 in flow sheet, 150-153 operating costs of, 449-450 power requirements for, 452 Flotation machines, 143-145, 147 capacity of, 144 determination of, 144-145 Flotation reagents, 31 consumption of, 33-34 for gold ores, 146 for silver 'ores, 148 Flotation tests, 29 FluoSolids, principles of, 165-166 FluoSolids furnace, 159, 165-168 installation cost of, 446 .Free gold, recovery of, 128-129 Fresnillo, treatment of manganese-silver ores at, 444-445 Furnaces, precipitate melting, 200-205 (See also Roasting)

Gold, free, recovery of, 128-129 production of, cost of, 448 economic aspects of, 11-15 world, 8-9, 12 retained in circuit, 207-208 revaluation of, 9 rusty, 129, 136, 181 in sulphides, 129 surface contamination of, 19-20 volatilization of, 273-274 Gold alloys, solubility of, 176 Gold chlorides, 274 Gold Coast, 386-395 Gold foil, dissolution of, 212-213 Gold Mines of Kalgoorli, Ltd., 410-414 Gold particles, size of, 228 Golden Cycle Corp., 327-330 classification at, 87-88, 90 roasting at, 165 sampling plant, 60 Golden Manitou Mines, Ltd., 320-321 "Golden :Mile," 398, 410 Golden Plateau, 415 Graphite ores, treatment of, 251 Silver-Dorfinan processes for, 251 at Timmons Ochali, 252 Gravity concentration, methods of, 128134 Gravity-concentration tests, 28-29 Grinding, control of, 236 Grinding analyses, 78

IT G Galena (galenite), 475 associated with silver, 437 Geita, 396 Gelatinous silica, 220 Getchell Mine, Inc., 330-331 carbon cyanidation at, 262-263, 265 Globe and Phoenix Gold Mining Co., Ltd., 385 Gold, dissolution of, as a corrosion process, 213 effect of oxygen on rate of, 212 physics of, 211 during washing, 126-127 distribution of, in tailings, 296

Hadsel mill, 53-55 Hallnor Mines, Ltd., 314-315 Hardinge classifier, 82-83 Hardinge "electric ear," 236 Hedley Mascot mill, 85 Hematite, roasting to, 162, 175 Hollinger mill, 315-318 concentrate treatment at, 155 costs at, 448 crushing plant at, 51 grinding control at, 236 precipitate treatment at, 204-205 Homestake Mining Co., 324-327 amalgamation at, 179-181 bullion parting at, 205-206

507

INDEX

Homestake Mining Co., classification at, 79 leaching at, 88-90 Merrill presses at, 111 Honduras, 342 Humphrey Spiral, 142 Hydrogen cyanide, 231, 254-261, 267 Hydroseparutors, 84

I Idaho-Maryland Mines Corp., 336, 338339 slime depression at, 149 India, 425-426 Infrasizer, 227-229 Ion exchange, 265-266

J Japan and Korea, 427 Jig beds, 134 Jig strokes and speeds, 134 Jigs, in mill circuit, 132-134 capacity of, 133-134 Juca Vieira Gold Mine, 347

Lake Shore Mines, Ltd., concentrate treatment at, 154 . pH control at, 237 roasting at, 159-162 sizing analysis at, 227-231 Lake View and Star, Ltd., 399-402 roasting at, 165 treatment costs at, 151 Leaching, of agglomerated slimes, 95 of clay ore, 92-95 Lead minerals, 475 effect of, in.cyanidation, 246-247 Lead salts, use of, 187, 210, 214, 219, 239, 243 Lime, 220 air slaked, 249-250 determination of, consumption of, 3638 free, determination of, 490 determination of, requirements of, 3536 use of, in regeneration, 244, 258-261 Lime plant in Nicaragua, 340 Loreto mill, 429-435 Lundberg, Dorr, and Wilson mill, 80 M

K Kalgoorlie gold fields, 398 Kalgurli are Treatment Company, 404409 Kenya, 396 Kerr-Addison Gold Mines, Ltd., 291-294 costs at, 448 Kelowna Exploration Co., 30D-302 slime separation at, 85 Kirkland Lake ore and treatment, 281 Kirkland Lake producers, 282 Knob Hill mill, 334-336 Kolar gold fields, 425-426 amalgamation practice at, 181 dewatering cones used at, 79 L

La India Companhia Minera, 337 La Luz Mines, Ltd., 337,340-341 Lake Shore Mines, Ltd., 303-307 calcine treatment at, 171-173

McClusky process, 444-445 Me Intyre Porcupine mill, 307-311 concentrate treatment at, 155 crushing plant at, 52-54 flotation in grinding circuit at, 152 pilot plant at, 150-151 MacLeod Cockshutt Gold Mines, Ltd., roasting costs at, 450 Magnetite, roasting to, 162,175 Manganese-silver ores, treatment of, 442-445 Marcasite, 239 Marievale Consolidated Mines, Ltd., 378-381 costs at, 451 power at, 453 Marlu Gold Mining Areas, Ltd., 392-394 desliming at, 84-85 Melting points of the elements, 458-459 Mensuration, 476 Mercury salts, use of, 242-243 Merrick weightometer, 236

508


Merrill-Crowe process, 185-193 Merrill filter press, 110-113, 189 Metallics, 27 Metals and their minerals, 474-475 Microns, definition of, 226 Microscopy in cyanidation, 17-20 Mill flow sheets, classification of, 279-281 Minerals, chemical and physical data on 474-475 Mochito mill, 342, 440-441 Modderfontein East, 139 Morro Velho mill, 347-350 control of cyanicides at, 243-244 Mother Lode ore, treatment of, 149-150 Mount Morgan mill, 415

N Negus Mines, Ltd., 307 Neptune Gold Mines, Ltd., 337, 340 New gold fields of Venezuela, 351 New Guinea, 427 New Occidental Gold Mines, 414-415 New Saza Mines, Ltd., 396 New South Wales, 414-415 New State Areas, 366-369 New York and Honduras Rosario Mining Co., 342, 438-442 New Zealand, 415 Nicaragua, 336-341 Nipissing mill, precipitation at, 198-199 Noranda Mines, Ltd., 321 conditioning at, 221 precipitate treatment at, 221

o Oliver filter, 118-119 Organic matter, 220 Osmiridium, 352 Oxidizers, chemical, 217 Oxygen in solution, altitude and saturation curves for, 215-216 determination of, 490-494 saturation of, 211-214

P Pachuca agitator, 103-104, 221, 240-241 Pachuea district, 429-433

Pamour Porcupine Mines, Ltd., 311-314 Pan-amalgamation process, 271-272 Passagem gold mine, 347 Patio process, 271 Peru, 350 pH determination, automatic recording of, 237,256 control of, 239, 241, 243-244, 256 in presence of lead salts, 247 indicators for, 21 instruments for, 21 theory of, 20-21 Philippine Islands, 421-423 Pickle Crowe Gold Mines, amalgamation at, 183-184 Porcupine ore and treatment, 283 Porcupine producers, 282 Porcupine United mill, 284-286 Portovelo mill, 345-346 Power requirements, for cyanidation, 451-453 for flotation, 452 Precipitate, clean up, 193-194 containing copper, 205, 207 filter press, 194 melting of, 200-201 treatment of, at Bibiani, 201-202 at Hollinger, 204-205 at N oranda, 202-203 Precipitation, aluminum dust, 195-199 on charcoal, 199-200 chemical control of, 186-189 chemistry of, 185, 197-199 economics of, 189 of gold in the laboratory, 42-43 at Kolar, 195 methods of, 184-185 ratio of, 235 on the Rand, 195 by sodium sulphide, 198-199 zinc dust, 187, 195 zinc shaving, 195-196 Presidio mine, 429 Preston East Dome Mines, Ltd., 286-288 power at, 451-452 Pre-aeration, 239-241 at Sub-Nigel, 375 Precoating, 184-185 Preliming in calcine treatment, 171 Primary slime, 30

509

INDEX

Protective alkalinity (see Alkalinity, control of; pH determination; etc.) determination of 489-490 Protective colloid, 149 Pulp consistency, formula for, 477-478 Pulp density control, 236 Pulp (slime) density, chart, 483 tables, 478-482 Punitaqui , 352 Pyrites, flotation of, 146, 150 gold bearing, 219, 238, 240-242 particle size of, 228 roasting of, 158-162 Pyrrhotite, 219, 238-243 roasting of, 159 at St. John del Rey, 348 at Sub-Nigel, 375

Q Queensland, 415 R

Radioactive tracers, 146 Rand (see Witwatersrand) Randfontein Estates Gold Mining Co., Ltd., 357-362 costs at, 450-451 power at, 452-453 precipitation at, 193 sand leaching at, 91 sorting at, 57 Reprecipitation, testing for, 40 Resende Mines, Ltd., 386 Revaluation of gold, 9 Rhodesia, southern, 385-386 Richards pulsator jig, 142 Rietfontein, calcine treatment at, 173174 Riffle, 130 Roasting, 158-176 chloridizing, 272-273 followed by leaching, 171-174 gold losses in, 170 opearting cost of, 449-450 in presence of lead salts, 247 at St. John del Rey, 350 with salt, 168-169 with soda ash, 169-170

Roasting, sulphate, 269 Roasting furnaces, 159-168 Roasting plants, cost of, 446 Roasting tests, 29 Rosario mill, 438-440 Russia, 425 Rusty gold, 129, 136, 181

S St. John del ~ey (see Morro Velho mill) Salsigne process, 241-242 Samples, minimum size of, 17 Sampling, ore and solution, 222 at Hollinger, 222-223 at Wright-Hargreaves, 222, 224-226 in ore testing, 16 theory of, 16-17 Sampling mills, 6G-61 Sand-slime separation, 90 Sand treatment, 87-95 at Homestake, 89 Screen opening vs. particle size, logarithmic plot, 482-483 Settlement, factors affecting rates of, 101-102 Settling rates for quartz spheres, 23 Settling tests, 43-44 Shafter, Texas, 429 Silver, metallic, 214 native, 428-429, 475 volatilization of, 273-274 world production of, 11, 12 Silver foil, dissolution of, in cyanide, 212 Silver minerals, 214, 219, 444,475 cyanidation and flotation characteristics of, 430-431 roasting of, 158 Silver sulphides, treatment of, 429-442 .Silver-Dorfinan process, 251-252 Sizing analyses, by air, 26 by beaker decantation, 22-25 of crusher house products, 54 by elutriation, 22 for mineral distribution, 226-229 at Pamour Porcupine, 313 plotted, 25 at Randfontein, 360 sieve, 21-22, 226

510


Sizing analyses, at West Rand Consolidated, 364 Slime, density of, chart, 483 tables, 478-482 depression of, 150 primary, removal of, 149-150 treatment of, 96-127 Sluice, 130 Smelter treatment, 153-157 SOo process at Kalgurli plant, 413-414 at Lake Shore, 159 for manganese-silver ores, 444-445 Sodium sulphide; 219 formation of, in solution, 239 leaching with, 265 precipitation by, 198-199 Solution, alkalini ty control of, 231-233 barren, in carbon cynnidntion, 263-264 color test for, 495-496 operating data in, 235 problem of, 234 sampling of, 225-226 change of, 39 clarification of, 184-186 cyanicides in, 233 cyclic use of, 40 de-aeration of, 186 grinding in, 39-40 pregnant, sampling of, 224-225 precipitation, inhibitors in, 223 testing of, 223 (See also Cyanide solution) Sorting, economics of, 58-59 in flow sheet, 57 at McKenzie Red Lake, 58 on the Rand, 57-58 by sink-float, 59-61 South Africa (see Witwatersrand) Standard mill, costs at, 450 Starch solutions, 44, 149-150 Steffensen machine, 144, 151 Stibnite ores, roasting of, 158, 249 treatment of, 249-250 Sub-Nigel Gold Mining Co., Ltd., 373376 Sulphides, concentration of, 128-130, 139 cyanide attack on, 242 flotation of, 143,151-152

SuI phides, gold-bearing, 128-130 roasting of, 158-170 in presence of lead, 247 treatment of, 153-157 Sulphotellurides, roasting of, 159, 165 treatment of, by bromocyanide, 267 (See also Tellurides) Sunshine mill, 437-439 Superpanner, 229-231

T Tables, bumping, 140-142 capacity of, 141-142 Tailings, plants for treatment of, in California, 331-333 pumps for, control of, 237 retreatrnent of, in Australia, 403-404 Tanganyika Territory, 396 'I'aquah and Abosso, 394-395 Tellurides, cyanidution of, 153 flotation of, 151, 152-153 particle size of, 228 solubility of, 250-251 treatment of, 154 (See also Suplhotellurides) Tellurides ores, treatment of, 250-251 Temperature-conversion formulas, flame and color scales, 476-477 Testing series, 38 Testing procedure, 26-45 Tetrahedrite, associated with silver, 437 Thickeners, Dorr, 96-97 traction, 97-99 tray, 97, 100 washing, 99-100 Thickening, general, 96 Thiocyanate, 219, 238-239 in barren solution, 235 determination of, 497-498 'I'hiosulphate, 219 in barren solutions, 235 'I'immons Ochali mill, 343-344 early table practice at, 140 graphitic ore treatment at, 252 Tonapah, Nevada, ore treatment at, 433437 Tramp-iron detector, 234 Tramp steel, removal of, 48 Trap, gold, hydraulic, 130-131

511

INDEX

U

Unit flotation cell, 130-132 United States, gold production in, 321324 flotation plants in, power consumption of, 452

v Vacuum, drainage, 91 Vacuum filters, 111-121 Van Dyk, 376-378 Vanners, 142 Vatukoula, Central mill at, 416-421 Venezuela, 350-351 Venterspost Gold Mining Company Limited, 381-385 Victoria Gold Dredging Co., 414 Volatilization of gold, 175-176

W Wanderers Consolidated Gold Mines, Ltd., 385

Washing, crusher feed, 289 flood, 120-121, 234 Washing tests, 29 Wedge furnace, 159, 162-164 Weighing of ore, 222 Wemco classifier, 75 West-Rand Consolidated, 362-366 West Springs mill, 369-370 Wheeler pan, 182 Wilfley table, 141 Witwatersrand, filter practice at, 121 general pract-ice at, 352-357 Wright-Hargreaves, 296-300 concentrate treatment at, 154 costs at, 299 sampling of ore at, 234 sampling of solution at, 222-226 Z

Zinc, determination of, 499-500 Zinc minerals, 475 effect of, in solution, 219, 245-246

Cyanidation and Concentration of Gold and Silver Ores-John v.N. Dorr

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