HIDROMETALURGIA.pdf

Pressure Hydrometallurgy Fathi Habashi Department of Mining, Metallurgy, and Materials Engineering Laval University, Quebec City, Canadá Fathi,[email protected]

© 2014 by Fathi Habashi. AII rights reserved Published by: Métallurgie Extractive Québec 800 Alain, #504, Sainte Foy, Québec Canadá GIX 4E7 Tel.: (418) 651-5774. E-mail: [email protected] http://pages.infinit.net/habashi http://works.bepress.com/fathi_habashi/

Distributed by:

To Nadia, Hani, and Hatem with love

Laval University Bookstore Zone Pavillon Maurice-Pollack, Cité Universitaire, Sainte-Foy, Québec City, Canadá G1V0B4 Tel.: (418) 656-2600, Fax: (418) 656-2665 E-mail: conseiller@,zone.ul.ca Dépot legal 2011 • Bibliothéque nationale du Québec, Montréal • National Library of Canadá, Ottawa ISBN 978-2-922686-22-7 Fathi Habashi, Pressure Hydrometallurgy Printed in Québec City by Les Copies de la Capitale, Inc, No part of this book may be reproduced or utilized in any form or by any means, electronic or mechanical, including photocopying and recording, or by any Information storage or retrieval system, without written permission by the publisher. Métallurgie Extractive Québec is a non-profít publisher registered in Québec City # 2240676462 devoted to diffusion of extractive metallurgy literature.

ui

Other Books by the Author

Published by Métallurgie Extractive Québec, Québec City and distributed by Laval University Bookstore except otherwise stated

Technical - F. Habashi, Principies of Extractive Metallurgy Volume 1: General Principies (422 pages), 1969 (reprinted 1980), (out of print), Gordon & Breach Volume 2: Hydrometallurgy (468 pages), 1970 (reprinted 1980), (out of print), Gordon & Breach Volume 3: Pyrometallurgy (493 pages), 1986 (reprinted 1992), (out of print), Gordon & Breach Volume 4: Amalgam and Electrometallurgy (380 pages), 1998 - F. Habashi (editor), Handbook of Extractive Metallurgy, 4 volumes, 2500 pages, WILEY-VCH, Weinheim, Germany, Also: John Wiley, 605 Third Avenue, New York, NY 10158-0012 - F. Habashi (editor), Alloys. Preparation, Properties, Applications, 312 pages, WILEY-VCH, Weinheim, Germany (out of print). Now available from Métallurgie Extractive Québec - F. Habashi, Metallurgical Chemistry, American Chemical Society, Washington, DC, Manual (279 pages), Audio Course (DVD, 5 hours playing time). Now available from Métallurgie Exfractive Québec - F. Habashi, Metals fi-om Ores. An Introduction to Extractive Metallurgy. 2003, 475 pages - F. Habashi, Pollution Problems in the Mineral and Metallurgical Industries, 1996. 150 pages - F. Habashi, Textbook of Hydrometallurgy, 2nd edition, 1999, 750 pages - F. Habashi, Textbook of Pyrometallurgy, 2002, 600 pages - F. Habashi, Kinetics of Metallurgical Processes, 1999, 376 pages - F. Habashi (editor), Progress in Extractive Metallurgy, Vol. 1, Gordon & Breach 1973, 239 pages (out of print). Now available from Métallurgie Exfractive Québec - F. Habashi, Chalcopyrite. Its Chemistry and Metallurgy. McGraw-Hill International Book Company 1978, 177, pages (out of print). Now available from Métallurgie Exfractive Québec iv

• F. Habashi, I. N. Beloglazov, and A. A. Galnbek (editors), International Symposium. Problems ofComplex Ores Utilization, Mineral Processing & Extractive Metallurgy. Special Issue, Gordon & Breach 1995, 280 pages (Out of Print). Now available from Métallurgie Extractive Québec - F. Habashi, Aluminum. History & Metallurgy, 2008, 160 pages - F. Habashi, Researches on Rare Earths. History and Technology, 2008, 125 pages - F. Habashi, Researches on Copper: History, Metallurgy, 2009, 400 pages - F. Habashi, Gold: History, Metallurgy, Culture, 2009, 277 pages - F. Habashi, Researches on Asbestos, 2011, 115 pages - F.Habashi, Mineral Processing for Nano-Scientists, 2010, 175 pages - F.Habashi, Extractive Metallurgy of Copper, 2012 Historical - F. Habashi (editor), Gellert's Métallurgie Chymistry, 1998, 500 pages - F. Habashi, D. Hendricker, C. Gignac, Mining and Metallurgy on Postage Stamps, 1999, 335 pages - F. Habashi, Extractive Metallurgy Today. Progress and Problems, 2000, 325 pages - F. Habashi, From Alchemy to Atomic Bombs, 2002, 350 pages - F. Habashi, Schools of Mines. The Beginnings of Mining and Metallurgical Education, 2003, 604 pages - F. Habashi, Ida Noddack (1896-1978). Personal Recollections on the Occasion of SOthAnniversary of the Discovery of Rhenium. 2005, 164 pages - F. Habashi, Readings in Historical Metallurgy, Volume 1- Changing Technology in Exfractive Metallurgy, 2006, 800 pages - F. Habashi, Postage Stamps: Metallurgy, Art, History, 2008, 125 pages - F. Habashi, The Copts ofEgypt, 2006, 92 pages - F. Habashi, Chemistry and Metallurgy in the Great Empires, 2009, 272 pages - F. Habashi, Science, Technology, and Society, 2009, 316 pages - F. Habashi, Aqua Science Through the Ages. An Illustrated History of Water, 2010, 166 pages - F. Habashi, Mining and Civilization. An Illustrated History, 2010, 510 pages - F. Habashi, Pyrite. History, Chemistry, Metallurgy,

Table of Contents

Preface After the second edition of Textbook of Hydrometallurgy was published in 1999, new developments have taken place that necessitated revising the book. Since no time was available to do this and since most of the development that took place was mainly in pressure hydrometallurgy, I decided to write this small book covering this topic only. It should be considered as a supplement to the Textbook to which the reader should refer to for back|ground Information. The book is in eight chapters as follows: [1] Historical introduction [2] Technology [3] General principies [4] Leaching processes in absence of oxygen [5] Leaching processes in presence of oxygen [6] Precipitation processes [7] Attempts to avoid autoclaves [8] Laboratory autoclaves and pilot plants

1 History of Pressure Hydrometallurgy 2 Technology 3 General Principies 4 Leaching Processes in Absence of Oxygen 5 Leaching Processes in Presence of Oxygen 6 Precipitation 7 Attempts to Avoid Autoclaves 8 Laboratory Autoclaves and Pilot Plants Index

The chapter on laboratory autoclaves from the Textbook have been revised and brought up to date and included in this book. I have already published most of the material in this book as articles in the technical press in which I referred to the original literature. Quebec City 2014

Fathi Habashi Fathi.Habashií íarul.ulaval.ca

vi

vn

1 15 45 59 89 153 181 197 231

1 History of Pressure Hydrometallurgy liilroduction [•;irly Work Russian research l'urther development Pressure leaching of Chemical industry Leaching of tungsten Precipitation under pressure revisited Ammonia leaching Acid leaching Work at the Mines Branch in Ottawa The plant at Fort Saskatchewan Nickel from pyrrhotite - pentlandite WorkatBerlin Recent Advances General References Books and conference proceedings Updates

ZnS concéntrate

1 2 3 4 5 6 6 7 8 9 9 10 11 11 12 13 13 13

INTRODUCTION The pioneer work on hydrothermal reactions of interest to metallurgy was conducted in Russia at the very beginning of the 20th century, mainly by Ipatieff and Bayer each working independently in Saint Petersburg. Gradually industrial applications took place first in the aluminum and later in the nickel industries. Today, the technology is well established in a large numbers of industries, e.g.. vni

Pressure Hydrometallurgy

uranium, copper, gold, tungsten, zinc, and titanium in addition to the aluminum and nickel.

EARLY WORK The first experiments that led to this technology were conducted in 1859 by the Russian chemist Nikolai Nikolayevitch Beketoff (1827-1911) (figure 1.1) while studying at the Sorbonne in Paris under Jean-Baptiste Dumas (1800-1884). Beketoff found that metallic silver can be precipitated from a silver nitrate solution when the latter is heated under hydrogen pressure. He also found that the initially neutral solution of AgNOg, turned acidic at the end of the test thus the reaction can be formulated as follows:

Chti¡>tir I - Introduction

Russian research This work was continued later in Saint Petersburg by Vladimir NikoInycvitch Ipatieff (1857-1952) (Figure 1.2) who in 1900 started a series of studies on numerous hydrothermal reactions under pressure. Among these was the precipitation of metáis and their compounds from aqueous solutions by hydrogen. He spent the first few years dcsigning a safe and reliable autoclave for these tests. Ipatieff's son joincd later in this research.

Figure 1.1 Nikolai Nikolayevitch Beketoff (1827-1911)

2Ag"+H2->2Ag + 2H" For his tests, Beketoff used a sealed glass tube containing the solution which acted as an autoclave. Hydrogen was introduced from a side compartment of the tube by the action of acid on zinc. At that time, of course, hydrogen was not available in cylinders - in fact gases were not yet liquified. Liquifaction of gases was introduced much later after Andrews experiments on the critical temperature and pressure of gases in 1869. He published his work in Compte Rendu de l'Académie de Science in Paris.

Figure 1.2 Vladimir Nikolayevitch IpatiefF (1857-1952)

Figure 1.3 Karl Josef Bayer (1847-1904)

At about that time, also in Saint Petersburg, Karl Josef Bayer (18471904) (Figure 1.3) an Austrian chemist working in a chemical factory to prepare aluminum hydroxide for mordanfing texfiles before dyeing, studied in 1892 the leaching of bauxite by NaOH at 170°C and pressure in an autoclave to obtain sodium aluminate solution from which puré AI(0H)3 would be precipitated by seeding at atmospheric pressure. A reactor dating from this period is shown in Figure 1.4 while today an autoclave is 7 m diameter and 40 m long (Figure 1.5).


I luipicr 1 - Introduction

uhcre M = Cu, Ni or Co. In the meantime, chemists started to use picssure vessels for a variety of reactions. Thus, Walther Nernst in Berlín discovered in 1907 that ammonia can be produced by the rcaclion of nitrogen and hydrogen at about 7000 kPa. Such a process was considered so impractical to him because of the "high pressure" involved that he did not care to patent it. It was Fritz Haber in 1913 who realized the technical importance of such reaction and built a bcnch scale unit that resulted in the development of one of the most outstanding achievements of the chemical industry - the ammonia synthesis. n

r^í

H

Figure 1.4 -Autoclave dating from the time of Bayer, about one meter long

Figure 1.5 - The largest autoclave is titanium ciad, 7 m diameter and 40 m long

Further development In 1903, M. Malzac in France patentad a process for leaching sulfides of copper, nickel, and cobalt by ammonia and air and recommended that high temperatures and pressures should be used for accelerating the rate:

Pressure leaching of ZnS Most metal sulfides are practically insoluble in water even at temperatures as high as 400°C. But, in the presence of oxygen, they are solubilized as sulfates. In 1927, Fredrick A. Henglein (1893-1968) (Figure 1.6) treated an aqueous suspensión of ZnS at 180°C with oxygen under 2000 kPa, converting it completely within 6 hours into zinc sulfate. The work was done in connecFigure 1.6 - Fredrick A. Henglein tion with purifying coke oven gas (1893-1968) from H2S. When scrubbing the gas with ZnSO^ solution, ZnS is precipitated. It is then regenerated by a hydrothermal reaction at 180°C (Figure 1.7): ZnS, + 2 0 , , ,^ZnSO^, , (s)

MS + 2O2+ nNH3-> [M(NH3)J2++ S O / -

2(aq)

4(aq)

Henglein found further that traces of copper sulfate or cadmium


sulfate in solution accelerated the dissolution. He attributed this to a catalytic effect: M2^ + Z n S ^ Z n 2 ^ + M S

Chapter 1 - Introduction

lite concentrates - a process that is now widely used in Russia and abroad. Maslenitsky was also the sénior author of the hodk Autoclave Processes in Nonferrous Metallurgy published in 1969 in Moscow by Metallurgiya Publishing House.

where M = Cu or Cd. ZnSO, solution

HzS- Free gas - ^ —

t i o:

Coke oven gas — • ZnS slurry

Oxygen

i

T

Pressure Leaching

Figure 1.7 - Purifying coke-oven gas containing HjS

Chemical industry

Figure 1.8 Ivan Nicolai Maslenitsky (1900-1972)

Precipitation under pressure revisited The Chemical industry, unlike the metallurgical industry, has been making an extensive use of pressure reactors. The hydrogenation of vegetable oils, the synthesis of methanol, the synthesis of ethanol, the Fischer-Tropsch reactions for organic synthesis, the Bergius process for hydrogenation of coals are only a few examples. Astonishingly high pressures have been used. Thus the ammonia synthesis by the Claude process utilizes 100 000 kPa and polyethylene synthesis utilizes 330 000 kPa pressures. Leaching of tungsten concéntrate In 1938 Ivan Nicolai Maslenitsky (1900-1972) (Figure 1.8) at the Leningrad Mining Institute (now St. Petersburg Mining University), developed an autoclave method for tungsten extraction from schee-

In 1946, the Chemical Construction Corporation, a subsidiary of American Cyanamid Company in New York City, which was in the business of building ammonia and nitric acid plants, had some problems in the removal of 00 impurity from synthesis gas, a mixture of hydrogen and nitrogen. This problem was given to Félix A. Schaufelberger (1821-2009) (Figure 1.9) a young gradúate from the Federal Institute of Technology in Zurich, Switzerland who had joined Cyanamid's Stamford Research Laboratories in Connecticut a year before. Towards the end of 1948, Schaufelberger succeeded in precipitating puré copper from sulfate solution by reduction with hydrogen in quantitative yield, liberating sulfuric acid for recycling in the leaching circuit. He had also prepared the first samples of nickel and of cobalt metal powder by this technique.

8



Acid leaching To supply cobalt to the Korean War efforts in 1950-53, the initial two projects at Calera in Utah, and at Fredericktown in Missouri, were rushed unduly without adequate piloting of process equipment. The final success, however, encouraged Maurice Dufour of Freeport Sulphur Company to contract for the development of an acid leach extraction process for laterite of Moa Bay in Cuba with Schaufelberger's flowsheet using high pressure leaching at 250°C. Work at the Mines Branch in Ottawa

Figure 1.9 Félix A. Schaufelberger (1821-2009)

Figure 1.10 FrankA. Forward (1902-1972)

Ammonia leaching It was also during this period that a new look at the oíd work mentioned above was considered by Canadian metallurgists. The ammoniacal leaching by Malzac's was applied by Frank A. Forward (1902-1972) (Figure 1.10) at the University of British Columbia in Vancouver on a laboratory scale for leaching a nickel-copper ore. Eldon Brown, president of Sherritt Gordon Limited, together with his consultant, Professor Forward went to Chemical Construction Corporation to discuss the design and engineering of a nickel extraction process which Forward had proposed, an oxidative leach of nickel sulfide in ammonia solution. The nickel powder prepared by Schaufelberger was presented to the visitors which convinced them immediately as a means of recovering nickel from solution and led to a cióse cooperation between the two companies.

In April 1956, all patents on this pressure leaching and pressure precipitation were transferred to Sherritt who used Forward's ammonia leaching combined with Schaufelberger's work to precipítate puré nickel from the solution obtained. The work was done at the Mines Branch in Ottawa before it was transferred to Fort Saskatchewan in Alberta: NÍS + 2O2+2NH3 [Ni(NH3)J2^ + H,

[Ni(NH3)2]2^ + S0,2NÍ + 2 N K

It was Vladimir N. Mackiw (1923-2001) (Figure 1.11) who made the discovery that copper could be precipitated from the leach solution as copper sulfide prior to nickel recovery, when the solution was boiled at atmospheric pressure, due to the presence of trithionate and thiosufate ions. This opened the way for direct nickel reduction from purified leach solution. Mackiw later developed a process for generating the fine seed particles for initiating the reduction process. It was observed that self nucleation sometimes occurred in solutions made up from pilot plant-derived nickel ammonium sulfate, but it never occurred in solutions made from puré salts. Details analyses revealed that the

10

Pressure Hydromeíallurgy

active ingredient was ferrous sulfate, present in trace amounts in the pilot plant solution, which led to the seed nucleation process. This process is still used in commercial operations, producing a significant amount of the world's puré nickel metal.

11


Nickel from pyrrhotite - pentlandite Work at the Mines Branch (now CANMET) by Kenneth W. Downes (1909-1996) (Figure 1.13) and his co-worker R.W. Bruce in 1955 demonstrated that pyrrhotite - pentlandite concéntrate could be treated in autoclaves in dilute acid at 120°C under oxygen pressure to get nickel in solution while Fe203 and elemental sulfur remain in the residue. The process was later applied by the Russians at Norilsk plant for nickel recovery.

• 1

^ P P W ^ ^ ^t^^^H

Figure 1.11 Vladimir N. Mackiw (1923-2001)

The plant at Fort Saskatchewan A by-product of this process was (NH4)2S04 which was marketed as a fertilizer. The process has been successfully in operation by Sherritt-Gordon since then, using a large number of autoclaves, and now used worldwide. From 1960 to 2001 all Canadian nickel coins were produced by this technology (Figure 1.12).

fj"-'-.

Figure 1.13 Kenneth W. Downes (1909-1996)

m^M Figure 1.14 Franz Pawlek (1903-1994)

Work at Berlin Extensive research on pressure hydrometallurgy was conducted at the Technical University in Berlin in 1960-1970 by Franz Pawlek (1903-1994) (Figure 1.14), his co-workers and others. Figure 1.12 - Canadian nickel coins produced by pressure hydrometallurgy from 1960 to 2001

12

Pressure

Hydrometallurgy

( hapter 1 - Introduction

GENERAL REFERENCES

RECENT ADVANCES The concept of pressure oxidation for treatment of refractory gold ore was developed by Sherritt in the 1980s, in collaboration with Homestake Mining Company (now Barrick Gold Corporation) for application at the McLaughlin project in California. About thirty pilot plant campaigns investigating this technology have been conducted by Sherritt since the 1980s. This led to successful commercialization at numerous gold operations in Canadá, Brazil, and Papua New Guinea. Large scale plants for the recovery of gold from refractory ores using high pressure technology went into operation recently in Finland, Russia, Dominican Republic, Brazil, USA, and Papua New Guinea. Hydrothermal reactions are now widely applied to treat directly zinc sulfide concentrates to get zinc in solution and elemental sulfur. Figure 1.15 gives a summary of these processes. Very large autoclaves 7 m diameter and 40 m long are used. Accessory units such as flash tanks and membrane pistón pumps are now standard equipment in a hydrothermal metallurgical plant. PRESSURE HYDROMETALLURGY

1

LEACHING 1

\ InAbsenceofOxyger Bauxite Kaolinite Ilmenite Laterite Antimondes Arsenides Pyrochlore Scheelite Wolframite

i

PRECIPITATION

+

{

1

i

InPresenceofOxygen

ByHj

ByS02

ByHaS

Sulfides

Nickel Cobalt

Copper

Nickel Cobalt

Disulfides Selenides Tellurides Uranium oxides

13

UO2

Figure 1.15 - Summary of hydrometallurgical processes

Books and conference proceedings l.N. Maslenitsky et a\., Autoclave Processes in Hydrometallurgy [inRussian], Metallurgia, Moscow 1969 S.S. Naboichenko et a l , Processing of Copper-Zinc and Zinc Concentrates Using Autoclaves [in Russian], Metallurgia, Moscow 1989 and the following at the end of Updates: F. Habashi, New Era in Pressure Hydrometallurgy Metall 68(1-2), 27-34 (2014) D.S. Flett and M. T. Anthony, Pressure Hydrometallurgy: A Review, Mineral Industry Research Organization, Lichfield, England 2000 M.J. Collins and V.G. Papangelakis, editors, Pressure Hydrometallurgy 2004, Canadian Institute of Mining, Metallurgy, and Petroleum, Montreal 2004 M.J. Collins, D. Filippou, J.R. Harlamovs, and E. Peek, editors, Pressure Hydrometallurgy '12,Canadian Institute of Mining, Metallurgy, and Petroleum, Montreal 2012

Updates F. Habashi, "Fállung ven Metallen und Metallverbindungen aus waBrigen Losungen durch Gase", Chemiker Zeitung (Heidelberg) 93 (21), 843855 (1969) F. Habashi, "Die Auflosung von Sulfidmineralien - Ihre theoretische Grundlage und technischen Anwendungen", Metall (Berlín) 24 (10), 1074-1082(1970) F. Habashi, "Pressure Hydrometallurgy: Key to Better and Nonpolluting Process", Part 1, Eng & Ming. J. 172 (2), 96-100 (1971), Part 2, ibid. 172 (5), 88-94 (1971) F. Habashi "Recent Advances in Pressure Hydrometallurgy", Proceedings International Conference on Advances in Chemical Metallurgy, Bombay 1979,1,18/1-18/34(1979) F. Habashi, "Recent Advances in Pressure Leaching Technology", paper S.4 in Proceedings First International Conference on Solvo-Thermal Reactions, Takamatsu, Japan 1994

14


F. Habashi, Industrial Autoclaves for Pressure leaching Technology", pp.6467 in Proceedings Second International Conference on Solvo-Thermal Reactions, Takamatsu, Japan 1994 F. Habashi, "Recent Advances in Pressure Leaching Technology", pp.l29139 in Volume 4 in International Mineral Processing Congress, edited by H. Hoberg and H. von Blottnitz, GMDB Gesellschaft fíir Bergbau, Metallurgie, Rohstoff-, und Umwelttechnik, Clausthal-Zellerfeld, Germany 1997 F. Habashi, "Hydrothermal Reactions of Sulfides and Disulfides", pp. 3949 in Proceedings Third International Symposium on Solvothermal & Hydrothermal Processes, Research Institute for Solvothermal Technology, Takamatsu, Kagawa, Japan 1997 F. Habashi, "Pressure Hydrometallurgy. Past, Present, and Future", pp.27-34 in Proceedings ofthe third International Conference on Hydrometallurgy, Kumming China, edited by Yang Xianwan, International Academic Publishers, Beijing, China 1998 F. Habashi, "Laboratory Autoclaves for Hydrometallurgical Research," pp. 411-418 in EPD 2000 edited by P. R. Taylor, TMS-AIME, Warrendale, PA 2000 F. Habashi, "Present Status of Hydrometallurgy Under Pressure" (in Russian), Komplexone Ispol'zovanie Mineral'nogo Sy'ya (1), 85-95 (2001) F. Habashi, "The Origin of Pressure Hydrometallurgy", pp.3-20 in Pressure Hydrometallurgy 2004, edited by M.J. Collins etal. published by Canadian Institute of Mining, Metallurgy, and Petroleum, Montreal 2004 F. Habashi, "Chalcopyrite -Atmospheric versus Pressure Leaching," Metall 61(5)303-307(2007) F. Habashi, "Chalcopyrite: Bioleaching versus Pressure Hydrometallurgy," pp. 17-22 in Proceedings International Conference: Metallurgy ofthe XXI Century. State and Development Strategy. Institute of Metallurgy and Mineral Beneficiation, Almaty, Kazakhstan 2006 F. Habashi, "New Era in Pressure Hydrometallurgy" Metall 68(1-2), 27-34 (2014)

Technology Hydrometallurgy Pressure Leaching Pressure Leaching Plant Autoclaves Materials of Lining of Pumping Agitation and mixing Heat transfer and economy Flash evaporators Slurry preheater Safety Mass transfer Design improvement Heat exchanges Transportation of autoclaves

construction autoclaves

15 17 18 19 31 32 36 38 38 38 39 40 41 41 41 41

HYDROMETALLURGY Generally, hydrometallurgy involves two distinct steps (Figure 2.1): • Selective dissolution of the metal valúes from an ore - a process known as leaching. • Selective recovery of the metal valúes from the solution, an operation that involves a precipiíation method. Sometimes a puriflcation/concentration operation is conducted prior to precipitation. These processes are aimed at obtaining a

16


puré and a concentrated solution from which the metal valúes can be precipitated effectively. The methods used are: adsorption on activated charcoal, sorption on ion exchange resins, and extraction by organic solvents. Ore Oxidant[

I

Leaching agent Le

Leaching

17

Chapter 2 - Technology

Leaching agent

Ore

Leaching

Solid-Liquid Separation

Valuable residue

Regeneration


Solid to waste

To disposal Figure 2.2 - Purification of ores by hydrometallurgy


Precipitant or electric current

Precipitation

Puré compounds

Metals

Figure 2.1 - General outline of hydrometallurgical processes

Sometimes hydrometallurgy is applied a chemical beneficiating method. In this case, the undesirable components of the raw material are leached away and the remaining solids are the valuable product that has to be processed further (Figure 2.2). For example, the treatment of ilmenite to produce synthetic rutile, the purification of cassiterite concentrates, etc.

PRESSURE LEACHING High-pressure leaching necessitates the use of pressure reactors (or autoclaves). One should distinguish between two types of pressure leaching: In absence ofair or oxygen. In this case the rate of leaching at ambient or modérate temperature is low and a temperature higher than the boiling point of the solution must be used. Consequently the reaction must be conducted in a closed vessel to prevent the escape of vapours. The pressure generated is the result of the vapour pressure of the solution. This method is mainly used for leaching bauxite, scheelite, ilmenite, and laterites. In presence ofair or oxygen. In this case leaching at ambient or modérate temperature is not possible unless air or oxygen is present as oxidizing agent. In both cases, it is the oxygen partial pressure that has the controlling factor on the rate of leaching. At a certain temperature the rate increases with increasing oxygen partial pressure. The use of oxygen instead

18


of air is more advantageous because for the same oxygen requirement the total pressure in the autoclave is low, and as a result the autoclave design will be less demanding, or decreased in size. This method is mainly used for leaching sulfides, selenides, tellurides, and arsenides.

19

< hapler 2 - Technology

hcat in the preheating exchanger of the incoming slurry and to make ihc hol slurry suitable for filtration. A simplified plant is shown in l'igure 2.5. FLAStíSTEAM

VW7T0 SCHUBGB?

When ammonium hydroxide is used as a leaching agent, the vapour pressure due to the volatility of NH3 should be taken into consideration (Figure 2.3). í'oiKCllUiíiC p!Cpai:!!101!

píüiiP

Figure 2.5 - A typical pressure leaching plant in which air or oxygen is used

Autoclaves According to their shape, autoclaves may be in form of cylinders vertically mounted or horizontally laid, spherical, or in form of a long horizontal tube. Method of agitation in an autoclave may be effected by injecting high-pressure steam, mechanically, or by rotating the whole autoclave.

40

80

120

160

Temperalure, C

Figure 2.3 - Vapour pressure of aqueous ammonia solutions

PRESSURE LEACHING PLANT A typical pressure leaching operation is shown in Figure 2.4. The central equipment in the plant is the autoclave. Its size has increased gradually from a modest 1 m long in 1889 to 40 m long and 7 m diameter - the largest to date. Other equipment include the high pressure membrane pump to forcé the slurry in the autoclave and the flash tank to decrease pressure of the exit slurry, to recover its

Vertical autoclaves are usually steam-agitated, some-times mechanically agitated, the horizontal are agitated mechanically by impellers and sometimes rotating, while the spherical are agitated by rotating the whole autoclave slowly around its horizontal axis. Some horizontal autoclaves are also agitated by rotation. Steam-agitated and rotating autoclaves have mínimum maintenance costs while autoclaves agitated by mechanical impellers are usually expensive to maintain because of the rotating shafts. Industrial autoclaves have volumes of 10 to 70 m^ and opérate at 2500-5000 kPa. Autoclaves are usually connected in series to achieve continuous operation. When laid horizontally they are usually mounted on a slope of about 8° to provide flow by gravity from one to the next.

~pl"

20


§

21

( hapter 2 - Technology

Nozzles for autoclaves are expensive and difficulties are encountcred in their design, particularly with a lead and brick lined vessel. it is therefore advisable to minimize their number even to the extent of múltiple Services per nozzle. It is also desirable to lócate as many of the nozzles as possible in the vapour phase; liquid-phase nozzles are subject to plugging. Dip-pipes extending into the liquid phase Irom vapour phase nozzles are used. Where liquid-phase nozzles are necessary, it is advantageous to provide means of back-flushing while in operation. Shield ~ - , i .

o c o CS

>

OJ

oa.

M o. 60

c o (L)

Opening for pulp inlet

Opening for síeam iniet

a> OH

I

r-i i>

Figure 2.6 -A 30 cubic meter vertical autoclave for leaching bauxites

Vertical, steam agitated autoclaves This is the simplest type and is used for leaching a material that requires no aeration, e.g., bauxite. An autoclave for this purpose is simply an insulated vessel capable of with-standing the operating pressure and is supplied with the necessary openings for introducing and discharging the pulp. These are usually fabricated from welded Steel cylinders with spherical ends. Diameters vary from 1.5 to 2 m and heights from 6 to 12 m. In the upper part are located apertures

22

. Pressure Hydrometallurgy

for admission of pulp, a manometer, safety valve, and a discharging pipe. Steam is fed through the bottom for heating and mixing. On their outside surface, the autoclaves have a layer of insulation. Such equipment is used mainly for leaching bauxite by NaOH usually at 140-150°C and 2500-3500 kPa. A typical design is shown in Figure 2.6. Autoclaves of similar design but with acid-resistant brick lining are used for leaching oxidized ores, e.g., laterites, by concentrated H2SO4 at 250°C and 4000 kPa. In the Moa plant, Cuba, the steel shell is first lined in the inside with 6 mm lead (Figure 2.7). Protective brick lining consists of acid-proof brick and of carbón brick. Under the usual operating conditions, the acid-resistant brick is subject to cracking. Carbón brick, on the other hand, does not. In this way, the carbón brick protects the acid-resistant brick from corrosión and erosión. All interior parts of the autoclave and also the connecting pipes are made of titanium. Stainless steel is used only in áreas where the temperature is lower than 100°C.

2lL

Air and HaSCí

Leve! ¡ndicator —4 ^ Pulp inpuí

r

> Puip discharge

mn^' Figure 2.7 - A 70 m^ vertical autoclave for leaching laterites at 250°C and 4000 kPa


23

Vertical, mechanically agitated autoclaves Sometimes these autoclaves are used, e.g., in leaching uranium ores (Figure 2.8).

Figure 2.8 - Vertical, mechanically agitated autoclave, 3 m diaraeter and 6 m high. Unmarked openings are for Instruments.

Horizontal autoclaves When oxygen is essential for conducting a reaction in an autoclave, it is necessary that a high rate of gas-liquid interaction is achieved. This is met in mechanically agitated autoclaves. These autoclaves are cylindrical vessels, horizontally laid, that are divided in the inside by partitions. In each chamber is an electrically driven turbine mixer from the top. The body of the autoclave can be lined with lead, with alloy steel, or with rubber. Feed slurry is pumped into one end of the autoclave and cascades from one compartment to the next. The average degree of filling is 65-70% static to allow sufficient disengagement space for the exhaust gases and to reduce the possibility of plugging the relief valve nozzle. The máximum diameter in Canadá is about 3.3 m to enable shipment of the finished vessel by rail (Figure 2.9).

24


25

i 'híipier 2 - Technology

rcagents are injected into the liquid phase. Slurry is discharged from the autoclave through a dip-pipe or by overflow through a nozzle located at the desired operating level. Agitator

Figure 2.9 -A horizontal autoclave 3.3 m diameter = 13.2 m long. Courtesy of Sherritt-Gordon Mines, Fort Saskatchewan, Alberta, Canadá

Larger diameter vessels may be employed but they have to be fabricated on the plant site. Compartment length is usually equal to four times the diameter; thus the length of a four-compartment 3.3 m diameter autoclave would be 13.2 m. At 65% filling, such an autoclave has a static operating volume of 77.3 m-'. Horizontal autoclaves are usually designed with four compartments, since too much volume will be lost in providing the static head required for the flow of slurry from one compartment to the next. A major advance took place recently in the design and construction of the horizontal autoclaves. Some of these reactors are 4.6 m diameter and 30 m long, divided into 5 partitions each equipped with an agitator. The autoclave is constructed of a 5 cm thick carbón steel shell lined with a 6 mm lead sheet, and two layers of acid resisting brick having a total thickness of 17 cm (Figures 2.10-2.13). Oxygen or air for leaching is always injected at a point under the bottom impeller to utilize the impeller for dispersión. Ammonia and other

V-r— - . ^ , — j ; - - ^ Agitator Manhole

~1- Slurry ^^,,^1^

Figure 2.10 -A large industrial autoclave 4.6 m diameter and 30 m long, lined with acid-resisting bricks, used for the oxidation of pyrite and arsenopyrite to libérate gold prior to cyanidation

Figure 2.11 - Mechanically agitated horizontal autoclave

26



27

The rotation and the impact of the balls result in breaking the impervious crust that forms on the ore particles thus accelerating their leaching. Solutions or slurries are introduced through a pipe in a hoUow trunion. Figures 2.14-2.18 show rotating cylindrical autoclave used for leaching tungsten and molybdenum concentrates at 225°C and 2500 kPa pressure, while Figure 2.16 shows a rotating spherical autoclave used for treating titanium ores.

Figure 2.12 - Horizontal autoclave for precipitating sulfides by H S

Figure 2.14 -A 10 cubic meter rotating autoclave for leaching tungsten and molybdenum concentrates

Figure 2.13 - Horizontal autoclave with cooling coils

Rotating autoclaves These may be cylindrical or spherical in shape, constructed of steel with the proper lining. They tura on heavy pivots at a speed of 8-15 rpm. Through one of these pivots the loading and unloading of the pulp and the admission of steam are carried out. Through the other, the driving and turning of the autoclave is accomplished. These autoclaves are usüally partially fiUed with steel balls; this type is used in cases where an insoluble reaction product is formed on the surface of mineral partióles that impedes the penetration of the leaching agent.

Figure 2.15 - Rotating horizontal autoclaves for leaching scheelite concéntrate with Na^COj solution at Bergla, Austria (Lurgi)

28


29

('hapier 2 - Technology

Figure 2.16 - Spherical rotating autoclaves installed in plant (United McGill, USA)

im}m>t^)m^/t//^M>umm^^^íS!miÉM^^^ Figure 2.18 - Cross section of a spherical, rotating autoclave

Figure 2.17 - Rotating autoclaves under construction (United McGill, USA)

Tube autoclaves In tube autoclaves, the slurry is pumped through one end and is discharged through the other. The system has been applied in Germany and in Czechoslovakia in the 1960s for the continuous leaching of bauxite. The slurry is pumped into an externally heated thick-walled tube about 30 cm in diameter and 30 to 50 m long (Figures 2.19 to 2.22). The major part of the heat is supplied by the slurry leaving the tube. Only at the extreme end of the tube, steam from an outside source is used for heating. The development of diaphragm-piston pumps that are able to reach 10,000-20,000 kPa made possible the application of this reactor. The system is characterized by extremely short residence time 2-3 minutes, high thermal efficiency, and low capital investment.

30


31


Flash tanks

^

nJ^

^

*^

*^

Steam/Sait

Heat exchanger

JL xc

tu

*«h«*riMtatriMMMdMtallM&Mh*iMHÍÍiMkM^ tSSSSS^^SSlSZISllSi^k

LE

aje Preheater

Outlet

-*" Condénsate

MF

Figure 2.21 -Tube autoclave Figure 2.19 - Tube autoclave unit (Lurgi).

Figure 2.22 - Hatch tube autoclave

Materials of construction Figure 2.20 - Tube autoclave made of titanium for leaching under oxygen pressure at the Vereinigte Aluminium Werke-Lipperwerk, Lünen, Germany (Lurgi).

Corrosive and erosive conditions are usually encountered in leaching processes and consequently, the proper selection of materials of construction is an important factor in the design. It should be noted.

32


however, that impurities in leach solution may drastically change the corrosion-resistant properties. For example, stainless steel is well suited to boiling nitric acid but deteriorates rapidly when the acid contains small amounts of chlorides or fluorides. A new titanium alloy containing niobium recently produced by Wash Chang (Ti-45Nb) is claimed to resist severe corrosión conditions and has been used for the pipes inside the autoclave. Lining of autoclaves When an autoclave is used in non-corrosive conditions for example in treating bauxite by NaOH then mild steel is used and is insulted from the outside by insulating material to prevent loss of heat. When HCI is used, usually the autoclaves are lined rubber as long as the temperature in 100-120°C and no oxygen is needed in the leaching. When the media is corrosive then autoclaves are lined either by acidresisting brick or by titanium cladding. Acid-resisting brick Prior to a newly lined vessel going into operation it is cured in an acid solution at a temperature around the boiling point. During this curing, a chemical reaction occurs in the brick that causes it to irreversibly swell, thereby, increasing the stresses and tightness of the lining system. For refractory lining, the advantages are: • Good corrosión resistance in sulfuric acid environment • Excellent abrasión resistance • Excellent resistance to oxidation and ignition The disadvantage, however, are: • Increased on-going rnaintenance costs • Requires larger vessels to accommodate refractory lining • Lower temperature limitations

('híipter 2 - Technology

33

Seiecting the right masonry materials for an autoclave lining sysicm is a challenging task considering the mechanical, thermal, and mechanical stresses encountered in pressure hydrometallurgical processes. The design of the brick linings has to take into account the almost non-elastic behavior of the brick inside a relatively highly clastic steel vessel. Bricks that have been successfuUy applied in autoclaves have a low AljOj content (~ 23%) and a high SÍO2 content (~ 70%). Bricks with higher AiÔj show a higher acid solubility. Depending on the abrasive characteristics of the slurry, ceramic bricks with an increased content of silicon carbide bricks (90%) SiC) can be used. Clay bonded SiC materials have been successfuUy installed. There are two types of fireclay masonry used in autoclaves: • Pressure vessel grade, is a less dense material that offers excellent acid resistance and good thermal shock resistance • Standard duty acid brick, is a dense material that offers excellent acid resistance but poor thermal shock resistance The mortar used for joining of the brick is a key component of a chemically resistant lining system. Resin mortars are formulated as two-component systems: liquid resins which act as a binder, and a powder containing inert filler and a catalyst. The catalyst causes the resin to cure when the two components are mixed prior to usage. Furan resin mortars have been used commercially for more than fifty years. They are obtained by the polymerization of furfuryl alcohol, co-polymerization of furfuryl alcohol and furfural, or by condensation of furfuryl alcohol and formaldehyde under acidic conditions. The inner filler is selected for its chemical resistance; carbón, silica, and barite powders are commonly used. Furan resins resist most acids and alkalies but not strong oxidizing agents. Henee they are only suitable for processes not using oxidants,

34

Pressure Hydrometallur^

e.g., at 150°C and in hydrochloric acid médium. Máximum service temperature ranges from 175 to 220°C. A disadvantage of these resins is their relatively high shrinkage during the curing process which may cause the brick to crack. The cracks, however, can be repaired easily by filling them with mortar resin before commissioning the vessel.

Figure 2.23 -Autoclave interior lining, 4.6 m diameter 25 m length, fíveagitator, 2-inch thick carbon-steel shell, 50.8 mm lead membrane, 2 layers of acid-resistant brick, 23 cm total thickness, capacity 415 m' (Barrick)

35

Chopicr 2 - Technology

TiUiniíim cladding Tilanium cladding Detaclad® Píate makes construction of autoclaves cconomically and technically viable. Titanium is required in all parís of the system to resist the hot sulfuric acid. In high-pressure urcas the equipment is fabricated from steel píate integrally ciad to titanium. The explosión welding process uses the energy of an explosión to créate a weld between metáis. The process is most commonly used to ciad steel with a thin layer of corrosion-resistant alloy metal, such as stainless steel, brass, nickel, silver, titanium, or zirconium. Although the explosión generales intense heat, there is not enough time for the heat to transfer to the metáis, so there is no significant increase in the temperature of the metáis. When two plates are being ciad, the mating surfaces of both metáis are ground fíat to achieve a smooth finish and prepare the surfaces Ibr the explosión. The plates are then ready to be assembled into the pack, which locks the plates into position. A small gap is left between the base metal and cladding metal. Next, explosive powder of exact formulation is evenly spread on top of the cladding píate. The explosión is detonated from one edge of the cladding píate and moves across the top of the pack at a uniform speed, which results in a high-pressure colusión of the metáis. The newly formed ciad is flattened out by a press (Figure 2.25 and 2.26). Pre-clad Assembly

Explosión Cladding Event Detonation Front

Space Between Plates

Figure 2.25 - Explosive cladding

Figure 2.24 - View inside of an autoclave during lining with acid resisting brick

Coltision Point

36


fuipter 2 - Technology

37

membrane, a pistón, and two ball valves. The space between the pistón and the membrane is filled with oil (Figure 2.28). When the pistón is in its downward stroke, the membrane expands outwards closing the lower valve and at the same time opening the upper valvc, thus forcing the slurry to move out, and vice versa, when the pistón is into upward stroke, the membrane moves inwards, opening the lower valve and closing the upper valve thus sucking the slurry in. The advantage of this pump is that the pistón does not come in contact with the slurry which can be abrasive.

Figure 2.26 -Autoclave manufactured from titanium-clad steel

For titanium ciad lining system, on the other hand, the advantages are: • Excellent corrosión resistance to oxidizing environment • Titanium can be in direct contact with process media resulting in smaller and lighter vessel • High temperatura limitation up to 300°C Disadvantages are:

Figure 2.27 - Typical installation of a high-pressure membrane pisten pump for pumping ore slurry into autoclaves

• Potential for ignition in enriched oxygen environment • Reduced abrasión resistance with low alloy grades • Susceptible to pitting and/or crevice corrosión in reducing environment

Pumping Transferring of solutions and slurries may be conducted by gravity flow when possible, but in most cases pumps are used. High-pressure membrane pistón pumps (Figure 2.27) are used for introducing pulps into autoclaves. The pump is equipped with a flexible rubber Figure 2.28 - Membrane pistón pump (section)

38

. Pressure Hydrometallurgy

Agitation and mixing Agitation and mixing oí solids in a solution may be conducted mechanically or pneumatically. In the first case an impeller causes the fluid motion while in the second case compressed air or high-pressure steam is used. Impellers are usually made to be about % to Vi the tank diameter, and if only one is on a shaft, it is placed no more than one impeller diameter from the bottom. When the impeller is in the center of the tank the motion is rotary and there is vortex formation. The liquid and solids are not forced sideways or vertically and as a result there is little mixing. This is especially the case for low pulp density slurries. To elimínate the formation of vortex two methods are commonly used:

39

('hapter 2 - Technology

dirccted towards the bottom of the tank where protective baffles are installed to minimize the erosión of the tank due to impact. This cquipment serves three purposes: • Decreasing the pressure and temperature of the slurry. • Recovery of heat in form of low-pressure steam. • Concentration of the solution as a result of the flash evaporation of water.

• Off center mounting of impeller either in axial or angular position. • Introducing baffles at the wall of the vessel. This produces an axial flow which is also necessary to oppose the settling of the particles. Baffles usually extend '/12 the tank diameter from the wall. Heat transfer and economy Heat transfer and economy is important only for pressure leaching processes where temperature as high as 250°C may be used. For endothermic reactions, e.g., leaching of bauxite, heat is supplied to the autoclave during the whole leaching period. On the other hand, for exothermic reactions, e.g., leaching of sulfides, heat is usually supplied only to initiate the reaction, and once this is accomplished, cooling will be necessary. The initial heating stage is usually done by direct steam injection. Flash evaporators These are large vertical tanks usually installed after an autoclave (Figures 2.29 and 2.30). The hot slurry is introduced through a tube

"y , Slurry outlet

Protection baffles

Figure 2.29 - Flash evaporators

Figure 2.30 - shows industrial installations for flash evaporators

Slurry preheater Slurries to be introduced in an autoclave are usually preheated by the steam generated in the flash evaporator. A typical design is shown in Figure 2.31.

40


('huilla V 2 - Technology

41

Mass transfer

J-Sluny

h is recommended that the introduction of oxygen in a pressure vesitcl to be below the impellers.

M B n h o i f t AccOKs "

Dcsign improvement DMtribution BaHls

Nii//,le design in flash tanks is of great importance since hydroiiiciallurgical slurries require flow control equipment (valves, and laiiks) made of expensive materials to withstand the abrasive and oflcn corrosive slurry. Ilcat exchanges

Marmol» Acovss -

st*«( sncll

Pr*-H**t*d Polp

Figure 2.31 - Slurry preheating by direct contact with steam

Safety When oxygen enriched air is used in leaching sulfide concentrates with ammonia the flammable conditions can be minimized by controlling the operating temperature, reducing the ammonia content in the solution, increasing the concentrations of nickel, copper, and zinc in solution, and keeping the oxygen content in the gas phase at lessthan 15.5%.

I leal exchanges are commonly used in a pressure hydrometallurgy planl. A self-cleaning fluidized bed heat exchanger is of good performance for laterite slurries. In this equipment the slurry is fed iipwards though a vertical shell and tube heat exchanger that has spccially designed inlet and outlet channels. Solid particles are also tcd at the inlet and these are carried through the tubes by upwards llow of slurry where they import a mild scraping effect on the wall ofthe heat exchange tubes, thereby removing any deposit at an early stagc of formation. These particles can be cut metal wire, glass, or ccramic balls with diameters varying from 1 to 5 mm. After passing through the outlet channel, the slurry and particles enter a separator where the particles disengage from the slurry and are returned to ihe inlet channel through an external down comer and re-circulated continuously. Transportation of autoclaves Most Most autoclaves have to be transported from manufacturing workshop to the mining site. This is a major engineering challenge bul it is done. For example, the autoclaves for Madagascar were iransported from Shanghai but those for the Dominican Republic

42



43

(Figure 2.33) were transported from Kuantan in Malaysia to Port of Samana. Each autoclave, weighing 750 tonnes, was lifted aboard the cargo ship for a four-week journey across the Pacific through the Panamá Canal to the Port of Samana. From there, the 120 km, 18day trip to the mine site required thorough surveying and temporary modiñcations to infra structure including bridges, roads, traffic signs and overhead obstructions. In the end, 27 bridges were reinforced with portable ramps or bypassed by temporary bridges.

Figure 2.33 -Transportation of an autoclave

Figure 2.32 - Transportation of an autoclave

Each autoclave was unloaded from the ship onto a pair of heavy-haul trailers, each with 22 sets of axles and 12 tires per row, and a 400tonnes-capacity turntable to allow trailer rotation under each end of the autoclave (Figures 2.33,234 and 2.35). Upon reaching the Pueblo Viejo mine, each autoclave was transferred to a self-propelled mobile trailer that was configured to comply with ground pressure limits and manoeuvrability constraints. Supported by auxiliary trucks for additional pulling, pushing, and braking capacity, the mobile trailer was able to manoeuvre the vessels cióse to their fináis site.

I

Figure 2.34 - Transportation of the autoclaves for Pueblo Viejo project [Hatch]

44

Pressure Hydrometallurgy'

General Principies Rccovery and Rate 45 Particle size 46 Concentration of leaching agent 46 Agitation 47 Pulp density 47 Temperatura 47 Effect of temperatura on the solubility of salts in water 48 Effect of temperatura on the solubility of gasas in water 48 The Boundry Layer 49 Diffusion-controUed processes 51 Chemically controllad processes 52 Intermediate-control processes 52 Aquaous oxidation of sulfides 53 General Principies of Precipitation 54 Nucleation and crystal growth 54 Co-precipitation 54 The precipítate 55 Disproportionation 56 Leaching Process 57

Figure 2.35 - Transportation of an autoclave

RECOVERY AND RATE In conducting a leaching process certain factors must be considered since they directly influence the cost of operation. For any leaching process, the percent recovery is a major concern. It is determined from a material balance based on the analysis of solids and solutions. The rate of a leaching process is foUowed by knowing the percent recovery as a function of time (Figure 3.1). The rate at any moment is the quantity of metal recovered per unit time. It is the slope of

46


the curve at that moment. It can be seen that at the beginning of the process the rate is high and then it decreases gradually with time. There-fore, a compromise should be made between the percent recovery and the residence time in the reactor to achieve máximum productivity. The rate of leaching depends on the following factors:

47

i/>/
ABttiition liicrcascd agitation usually increases the rate of leaching. But, this «(tilín may be a costly item due to increased capital cost of the agitator «lul opcrating cost due to the power consumed to effect agitation. I*ul|) density

100

Rate = slope

[Rate = slope

Katc of leaching increases with decreasing pulp density, i.e., when Itirgc volume of leaching agent is added to a small weight of solids. Hut this will result in dilute solutions that will make handling and recovery more difficult. High pulp density results in concentrated Holulions, but, on the other hand, is usually associated with high erosión of equipment, e.g., pumps, agitators, etc. Tcmperature

Time

Figure 3.1 - Recovery and rate of leaching

Particle size

Ralo of leaching increases with increased temperature (Figure 3.2). However, increased temperature may result in increased dissolution of other minerals thus causing increased reagent consumption and increased impurities in solution.

The rate increases with decreased particle size of soHds, i.e., increased grinding since the smaller the particles, the larger is the surface área per unit weight. However, increased grinding is not only a costly item, but also may cause a filtration problem. Extremely fine particles are sometimes difficult to settle and to filter. Concentration of leaching agent Increased concentration of leaching agent increases the rate of leaching. But, it may also cause the dissolution of undesirable minerals thus leading to increased reagent consumption and increased impurities in solution.

Temperature

Figure 3.2 - Increased rate of leaching with increased temperature

48

49

( hapter 3 - General Principies


Effect of temperature on the solubility of salts in water Most salts when they dissolve in water they absorb heat, consequently according to van't Hoff's equation:

S,_ AH °^ S. 2.303R the solubility increases with increased temperature up to 120-150°C. Beyond this temperature range it was found that the solubility decreases resulting in precipitation because of the vigorous thermal vibration of the water of hydration (Figure 3.3). 80 70 O £60

r

1

1

T-

¡

50 100 150 200 250 300 350 400 Temperature, °C

Figure 3.4 - Solubility of oxygen at high temperature

:

- /

'"> \ •^' \ \ _/ / 8 50 / ''' \\ \v \ - / / \\ \znSO4 'S)40 // \\ \ O)

\ \\CoS04\ \

= 30 cg20

J 1

^\ \-

10 _

FeSOÍ,\

\

\

\

! 50 100 150 200 250 300 Temperature, C

50 100 150 200 250 300 Temperature, C

Figure 3.3 - Solubility of salts at high temperature

Effect of temperature on the solubility of gases in water

50 100 150 200 250 300 350 400 Temperature, °C

Figure 3.5 - Solubility of hydrogen at high temperature

Contrary to the dissolution of solids, the dissolution of gases in water is accompanied by the liberation of a small amount of heat. Henee the solubility decreases with increased temperature. However, above the boiling point of water, the solubility of gases increases (Figures 3.4 and 3.5).

THE BOUNDRY LAYER A solid in contact with a liquid is covered by a stagnant film of liquid about 0.03 mm in thickness called the Nernst bound-

J]


50

ary layer after its discoverer Walther Nernst. Its existence is manifested in a stream line flow where the velocity of a liquid in a pipe is máximum at the center and gradually decreases to zero at the inside walls. This concept was applied to explain the dissolution of a solid in water. When a solid was agitated in water and the solution analyzed at intervals, it was found that the rate of increase of the solute concentration followed by the equation: ^ = k(C3-C)

< hofter 3 - General Principies

51

Thus, the rate constant k was identified by =0A/6V. This explained why the rate of dissolution increased with increased agitation because under these conditions, the thickness of the boundary layer decreases henee the increased rate of dissolution. The rate equation for physical processes was later extended to Chemical processes. It has been supposed that the reacting species must diffuse through the boundary layer before reacting. The rate equation would then be:

-f = k(C-C,) where C is the concentration of the solute at time t, Cg is the solubility of the compound in water at the experimental temperature, and k is the velocity constant. It was suggested that a saturated layer is rapidly formed at the interface and that the observed velocity is the rate at which the solvated molecules diffuse from this layer into the bulk of the solution. On applying Fick's law of diffusion to this process, then at a constant volume: Â dt

(C,-C)

where n is the number of species diffusing in unit time, 2) is the diffusion coefficient, A is the surface área of the solid, and 5 is the thickness of the saturated layer adhering to the surface of the solid, i.e., the boundary layer. Since:

where k is the rate constant, C is the concentration in the bulk of the solution, and C. is the concentration of the reactant at the interface; the negative sign indicates a decrease in concentration. The interaction between a solid and a leaching agent therefore takes place through the following steps: 1. Diffusion of reacting species to the interface 2. Adsorption at the interface 3. Reaction at the interface 4. Desorption of the products 5. Diffusion of the products from the interface Any of these may be rate-controlling depending on its relative speed with respect to the others. Diffusion-controIIed processes

C=

V

where V is the volume of solution, the above equation becomes: f = ^ ( C 3 - C ) = k(C3-C)

This is the case when the rate of chemical reaction at the interface is much faster than the rate of diffusion of reactants to the interface, resulting in C. = O

-Tf"\ 52


Rate= Â(C-C¡) = kÂC These processes are characterized by: • A strong dependence on the speed of agitation since agitation decreases the thickness of the boundary layer. • SUght dependence on temperature since the rate of diffusion is only slightly influenced by temperature.

53

('hapter 3 - General Principies

Rate =

k.k 12 k^-Hk,

AC = kAC

wherek = k^k2/(k^+k2). This represents the general case of leaching processes. If k^«k2, then k = k^ = ^ / 5 , i.e., the process is diffusion-controlled. If k2 « k^, then k = kj, i.e., the process is chemically-controlled. Aqueous oxidation of sulfídes Sulfide minerals are insoluble in water even at high temperature. They can be solubilized, however, by a variety of methods:

Chemically controlled processes This is the case when the rate of chemical reaction is much slower than the rate of diffusion henee it determines the observed rate: Rate = V^NZ.

These processes are characterized by independence on the speed of agitation because diffusion does not play an important role, and strongly dependent on temperature since the rate of chemical reaction increases exponentially with temperature.

• In absence of oxidizing agents. Some sulfides dissolve in acids forming H2S, others dissolve in basic médium forming sulfide ion. • In presence of oxidizing agents. Elemental sulfur usually forms but because of its instability in neutral or basic media it oxidizes to sulfates. In acid médium, however, there is a narrow región where elemental sulfur can form (Figure 3.6). The diagram is valid in the temperature región 25 to 150°C. Above 150°C, the narrow stability región of elemental sulfur disappears and no sulfur can form.

Intermediate-control processes HSOÍ c

This is the case when both rates are of the same magnitude, i.e., when a concentration gradient is formed across the boundary layer, but C. 7^ 0:

-

o \ ^ ^''"-v,..^^^

\j

Elementa! sulfur „

o-

^'"^C^x

Therefore,

'-

^ C

-

en c

Ü 3 T3 a>

°^

C.=

S05"

H2S

Rate = kÂ(C - C¡) = kÂC.

S2>Basic

Acid 7 PH

Figure 3.6 - Potential-pH diagram for sulfides at 100°C

54

Pressure Hydrometallurg)

GENERAL PRINCIPLES OF PRECIPITATION Nucleation and crystal growth Precipitation involves two steps: Nucleation and crystal growth. Factors favoring increased rate of nucleation are: concentrated solution, high speed of agitation, and the presence of finely divided solid in the solution which act as nucleating agent. If the rate of nucleation is high, the precipítate will be finely divided. On the other hand, if the rate of nucleation is slow, the precipítate will have large particle size. As a result, finely divided precipitates are obtained from concentrated Solutions and coarse precipitates are obtained from dilute Solutions. Rate of precipitation may also decrease as a result of the presence of certain metal ions or organic compounds in solution. Usually, precipitation is carried out under strict conditions to achieve the necessary separation from the other constituents. In general, the following factors control a precipitation process: • Precipitation should be conducted within a certain pH range, since most precipitates are re-dissolved outside this range. • Precipitation should be conducted at an optimum temperature since most precipitates are more soluble in hot than in cold solutions. Co-precipítation Co-precipitation is the contamination of a precipítate by substances that are normally soluble under the conditions of the precipitation. It can be an inconvenience due to the difficulty of obtaining a desired puré precipítate. It may result because of the following reasons: Occlusion This results when trace impurities at the surface of a crystal become mechanically trapped as the crystal layers are deposited. This type

•\Mvi >' General Principies

55

<»l precipítate may be purified by re-pulping in water, filtration, and <*<>shing. Adwrption Ihi.s is especially important when a gelatinous precipítate with a Inrgc surface área is formed. Trace impurities are adsorbed on the •urfacc. In some cases, the forces of adsorption are weak and the impurities may be desorbed by hot water. In others, surface reacIjons may take place and removing by washing is not possible. For •xample, when ferric hydroxide is precipitated, is usually adsorbs a number of other ions. h'ormation of solid solution This takes place when the ionic radii are nearly the same thus an impurity ion can substitute an ion of the macro component in the crystal lattice. Substitution takes place at random and therefore the impurity ion forms a solid solution with the macro component. In this type, it is not possible to remove the impurity by washing. The precipítate l'he particle size and form of a precipítate depend upon the conditions under which it has been formed. Particle size A freshly formed precipítate is sometimes described as amorphous or gelatinous because of the difficulty of separating it by filtration. The particle size of such precipítate may be small and ímperfect crystals due to the fast rate of precipitation. Precipitates usually undergo contínuous re-crystallizations when left in contact with the mother líquor; the process is called "aging". Ions continuously go into solution and redeposit on the surface of the solid and a state of equilibrium is achieved. The process is accelerated by heating. For example, a freshly precipitated hydroxide is difficult to filter and wash. On heating before filtration, it is converted into a non-gelati-

56


nous form rendering it easy to handle. Freshly prepared Fe(0H)3 undergoes a process of condensation, whereby HjO molecules are split, until finally water-free Fe203 is formed: 2Fe(OH)3

57

hiipter 3 - General Principies

LEACHING PROCESS Ixaching processes at high temperature and pressure may be conducted in absence or in presence of oxygen (Figure 3.7).

Fep3 -H 3 H P

Crystal form The crystal form of a precipítate may depend on the médium of precipitation. Por example, |3-FeOOH is formed when ferric ion is precipitated from chloride or fluoride médium while a- and yFeOOH are precipitated from sulfate, nitrate, and bromide médium. Aluminum hydroxide precipitated by neutralizing aluminum ion from acid médium is gelatinous, difficult to filter and wash free from impurities, while that precipitated from alkaline médium, e.g., sodium alumínate solution, is crystalline and readily filtered and washed. Disproportionation This involves the simultaneous oxidation and reduction of an ion; application in hydrometallurgy is limited to cuprous ion which can lose an electrón to become cupric, and gain an electrón to become elemental copper simultaneously: Oxidation: Cu* -^ Cu^* + e" Reduction: Cu* + e " ^ Cu Overall reaction: 2 C u * ^ Cu + Cu^* It can be seen from the overall equation that half the cuprous ion is precipitated as a metal and the other half remains in solution but at a higher valency state. In acid médium, this reaction is usually slow at ambient conditions; a temperature of 150 to 180°C is needed to have appreciable transformation.

1

LEACHING 1

In Absence of Oxygen Bauxite Kaolinite Ilmenite Laterite Antimondes Arsenides Pyrochlore Scheelite Wolframite

1

In Presence of Oxygen Sulfides Disulfides Selenides Tellurides Uranium oxides

Figure 3.7 - Pressure leaching processes

Leaching Processes in Absence of Oxygen líauxite Introduction Bayer Process Bauxite, Clay, Shale, Anorthite, Nepheline, Fly ash Kaolinite Introduction Pressure leaching of clay Latentes Introduction Latente in Cuba Laterite in Australia Laterite in Papua New Guinea Laterite in Madagascar Laterite in Turkey Ilmenite Introduction Rutile, Sorelslag, Synthetic rutile Methods of treating ilmenite Wolframite and Scheelite Introduction Alkaline leaching Introduction Treatment of pyrochlore Arsenides And Antimodes Introduction Purification of chalcopyrite concéntrate

60 60 62 68 68 68 70 71 71 71 73 76 78 80 81 81 82 82 84 84 84 85 86 87 87 87

60


61

\Jîj
Table 4.2 - Composition of typical bauxites

BAUXITE

%

%

Introduction Bauxite, named after the village Les Beaux near Marseille in southern France where it was first discovered, is not a mineral, but designates various kinds of aluminum ores consisting mainly of aluminum hydroxide. Three aluminum hydroxide minerals occur in bauxite: gibbsite, bóhmite, and diaspore. They differ considerabjy in theír physical properties, as shown in Table 4.1 bauxite deposit consists mainly of either one of these types, although cases are known when mixed hydroxides are present in one ore. Bauxites vary in color from cream to dark brown when the iron content is high. Table 4.2 shows the composition of typical bauxite. The main occurrences of bauxites are in Jamaica, Suriname, Ghana, Sierra Leone, Australia, Russia, and Hungary.

0.01

AÍ203

40-60

SÍ02

1-6

K2O

0.01

2-25

P2O5

0.02-0.4

1-5

V2O5

0.01-0.1

FejOB TIO3

CaO + MgO Loss on ignition

0.2-0.6 10-30

GaÔj

LnÔj

F

0.01 0.01-0.05

fíic treatment of bauxite to produce puré AI(0H)3 from which puré «ilumina is obtained by the Bayer Process using sodium hydroxide solution is the oldest and the largest pressure leaching operation in Icrms of the tonnage of raw material treated (Figure 4.1).

Table 4.1 -Aluminum minerals in bauxite Gibbsite Bóhmite

Diaspore

y-AI(0H)3

y-AIOOH

a-AIOOH

1 :3

1 :1

1 :1

monoclinic

orthorhombic

orthorhombic 6.5-7

(Hydrargillite) Formula AljOj : HjO Crystal system Hardness (Moh)

2.5-3,5

3.5-4

Specific gravtty

2.42

3.01

3.44

Refractive Índex

1.568

1.649

1.702

Temperature o f rapid d e h y d r a t i o n Product of dehydration

150°C

350°C

450°C

Z-AI2O3

y-AIÔs

a-Al203

128

54

insoluble

Year

Solubility in 100 g/L NajO solution at 125°C;g/LAl203

Figure 4.1 - Toimage of bauxite treated annually

Aluminum minerals in bauxite are soluble in dilute H2SO4 but this acid is not used on large scale for the following reasons:

62


63

\(. hapter 4 - Leaching Processes in Absence ofOxygen

BAUXITE

• Iron minerals and to some extent titanium minerals are also soluble; this will lead to an excessive reagent consumption and solution purification problem later. • AI(0H)3 precipitated from acid solutions is gelatinous and difficult to filter and wash.

T Grushing

I Washing

I

NaOH

Drying

1

NaOH make-up

Grinding

Acid leaching is used only on a small scale to produce aluminum sulfate needed for water treatment.

Leaching Settüng

Bayer Process The use of sodium hydroxide to leach bauxite was invented in 1892 by Karl Josef Bayer as a process for obtaining puré aluminum hydroxide which can be calcined to puré AijOg suitable for processing to metal. About 90 million tons of bauxite are treated annually by this process. About 2 tons bauxite yield 1 ton AlÔg from which 0.5 ton aluminum is produced. Also 2 tons bauxite produce 1 ton waste minerals called red mud. Crushed bauxite is usually washed to remove fine particles of clay, dried in a rotary kiln then ground to 60-100 mesh; the drying process is essential to facilítate grinding. Drying temperature should be less than the temperature of dehydratation of aluminum hydroxides otherwise the solubility will be impaired. Figure 4.2 shows a flowsheet of the process and Figures 4.3 and 4.4 show an operating plant. The reactions in leaching are the following:

Precipitatíon >•] Washing | • >\

Calcinalion

Evaporaüon Cenirifuge

" Soiid impuhties

Recycle

Figure 4.2 - Flowsheet of the Bayer Process for the production of AlÔ^ from bauxite ^Hv!^

H

t*»__«i*

H i^ ^nl ' H ^ ív . ^ ' 1

'JHh ^^

i

^29HBI

AI(0H)3+0H-^[AI(0H)J*

AIOOH + O H - ^ [AiOÍOHy-

Puré | - > - AiaOa

'

' - ^ j '

'

1 '

II

-%

:Á., 1• " •<•

— "-^'^î^PE ..

- i

'-a^^^T^'

'^^^^"'Z- "' ~^¿*-^T' ">''-

l^^

1 ] 1 jrtíÉSf^ • ^ P?^«^«^l^^: •_:.} i '-'4

- ^ • ^

> ' - -^ -Jif^SSí^S^^ ^

1

;

—' "J^-—^S^^^'íl

ESP**

'" ^ ^ p ^

T - - ' ^ l " ^

P^^^\^^DB93|

h¿M

Figure 4.3 - Bauxite treatment plant by the Bayer Process in Kwinana near Perth in Australia produces 1 250 000 tons/year alumina (ALCOA).

64


1»

ST£A»5.5H» .1.

C ,

SIEÍ«,lílI«

ll'll!

a

A «

r 1

Á— 1^

1

I

i.

A,

V^

O

k

•-•»

•

^

r

r

! '^ r

-i i i

3UFr£R VESSEl

« A I POLP

1 \ Pisto»

1

p»«?

1

.

-ODO——0

'

'

V'

y

^

,^

LJ 1

••

l£A0HE3

mi^

r m%

f-

- AjrOCLAVtS

S£?ABAIOfiS

A

Figure 4.4 - Continuous leaching of bauxite

Leaching is usually carried out in mild steel autoclaves, with direct steam injection for heating and agitation. Operating conditions depend on the type of minerals in the bauxite. Bauxites containing gibbsite are leached at a lower temperature, with lower NaOH concentration, and for a shorter time than those containing bohmite and diaspore as shown in Table 4.3 The more concentrated the NaOH, the faster the rate; however, highly concentrated solutions will require excessive dilution in the later stage of precipitation, which presents difficulties in handling and filtration. Therefore, there must be an optimum concentration which compromises between digestión time and subsequent operations. Leaching time could be shortened to 3-4 minutes if the process is conducted in tube autoclaves at 330°C and 25 000 kPa; also, the settling properties of the mud are improved. Table 4.3 -Typical leaching conditions of bauxite with NaOH Ore

Temp.,°C

Pressure, kPa

Gibbsite

140

400

140

1

Bohmite

180

800

350-600

2-4

Diaspore

180

800

NaOH,g/L

350-600

Time, h

lo a leach solution containing 200-250 g/L Na2C03, enough lime is «ddcd such that the solution contains about 140 g/L NaOH. Leaching is carried out at 140°C for about one hour. This method applies only to gibbsite, because in the case of bohmite or diaspore a sodium hydroxide concentration above that which can be obtained directly l'rom Na2C03 and lime is required. So it is necessary to prepare NaOH separately and concéntrate it by evaporation for use in leaching. The behavior ofimpurities Depending on their chemical properties, impurities in the ore may be found either in solution, or in the insoluble residue. Impurities ihat go into solution are either re-deposited during a later operation, or remain in the mother liquor during crystallization of aluminum hydroxide. However, due to recycling, they will accumulate in solution and would contamínate the product. Certain tolerable levéis are therefore maintained by bleeding the solution at regular intervals. Table 4.4 shows the distribution ofimpurities during the processing of bauxite, and Table 4.5 shows the composition of the Icach solution. Essentially, the Bayer Process eliminates the three major impurities in bauxite Fe203, SiO^, and T¡02. Calcium and magnesium are mainly present as dolomite and are not dissolved. Apatite is partially soluble and is usually eliminated during leaching by adding CaíOH)^. Table 4.4 - Distribution ofimpurities during the processing of bauxite Bauxite %

2-4

Sometimes NaOH is formed in situ in the autoclaves by adding sodium carbonate and calcium hydroxide:

65_

Na2C03+ Ca(0H)2^ 2NaOH + CaCOg

SIEíH

-*-—rn

vT.

ijuer 4 - Leaching Processes in Absence ofOxygen

Red mud (driedat105°C)%

Calcined AI2O3 %

AI203

57.8

14.0

99.55

SÍ02

3.5

7,6

0.05

Fe,03

24.3

57.6

0.04

TiOj

2.5

5.7

—

NaÔ

—

7.4

0.18

12.5

7.7

0.18

Loss on ignition

66

Pressure Hydrometallurgy:

Table 4.5 - Composition of typical aluminate solution g/100 g free N a p AI203

32.80"

ci-

1.00

CaO

0.12

CO,

3.74

Fe,03

0.05

F-

0.03

GajOl NajO

0.22

MgO

0.17

100.00

P2OS

0.90

SÍO2

0.60

SO3

0.07

TÍO2

Trace

V2O5

0.45

* In the precipitation step, about 2/3 of the aluminum in solution is precipitated.

Organic matter Most bauxites contain about 0.1 % organic matter. During digestión, some of this material is dissolved, causing the liquors to darken, while the remaining part is degraded and oxidized to oxalates. Some of the organic matter is therefore responsible for NaOH losses. Their presence may also cause liquors to froth, or it may interfere with the subsequent process of hydroxide precipitation, or it may color the hydroxide. It may be largely eliminated during digestión by adding an oxidizing agent, e.g., Mn02. In some cases sodium oxalate is crystallized and removed. It should be noted, however, that the organic matter in the liquor may come from the flocculants added to assist the mud to settle. Iron Iron occurs in bauxite mainly as hematite, Fe203, and is not attacked by the caustic leaching. Thus the residue remaining after leaching has a high percentage of iron oxides, and therefore has a red color. That is why this residue is usually referred to as red mud. However, some ores contain ferrous iron in form of siderite, FeCOg. This is attacked by NaOH, forming colloidal ferrous hydroxide: FeC03+ 2 0 H - ^ Fe(OH). + CO,^-

iplur 4 - Leaching Processes in Absence ofOxygen

67

.vliich is difficult to settle. It would therefore be advantageous to .•\idize ferrous minerals to ferric during leaching. Silicon Silicon occurs as quartz, SiO^, or as clays, e.g., kaolinite, AI^(OH)^(SÍ205). Quartz is insoluble in NaOH under the condilioiis of leaching but the silicates are soluble. During digestión, «ihca that goes into solution combines with alumina and sodium hydroxide forming insoluble hydrated aluminosilicates such as 2Na20'2Al203«3Si02*2H20, which are carried away in the red mud, ihus causing losses. About 1 kg of NaOH is lost for each kilogram of soluble silica in bauxite. Although most of the soluble silica in bauxite is precipitated during digestión by forming sodium aluminosilicate, small amounts may still be found in solution, especially when concentrated NaOH solution is used. To precipítate the silica completely, addition of CaO is recommended, since insoluble calcium silicate can be formed. Lime addition during digestión has a further advantage: any Na2C03 present in the solution due to absorption of CO2 from the atmosphere, and which has no dissolving action on bauxite, will be converted to NaOH. Gallium Gallium occurs in most bauxites, and is highest in French bauxites (0.0-0.05 % 08203). It dissolves completely during extraction. Recycling of NaOH in the process results in gallium enrichment up to 0.5 g/L. Such liquors are therefore an important source of gallium, from which it can be recovered, e.g., by solvent extraction or electrolysis using a mercury cathode. If, however, gallium is not recovered and is left to build up in the leach liquor, it will reach a certain concentration beyond which it will be partially deposited, together with aluminum hydroxide during precipitation, thus causing contamination.

68


Vanadium Vanadium in bauxite is partly soluble during digestión. In some ores it is precipitated during evaporating the leach solution as complex salts such as 2Na3V04*NaF»19H20. This is especially the case for ores containing fluorine since fluorides are dissolved during leaching. Sometimes these precipitate to form a hard scale in the evaporators which interferes with heat transfer. In other ores vanadium builds up in the recycle NaOH to a concentration of about 0.5 g/L VjOg and is recovered.

KAOLINITE Introduction Kaolinite is the most important mineral in clays. North America produces about 50% of the world's aluminum, yet must import more than 90% of the raw material needed, although there is abundant domestic resources of aluminum-bearing silicates raw materials such as clay, shale, anorthosite, nepheline, and fly ash from power plants (Table 4.6). That is why there is extensive research underway to recover alumina from these non-bauxitic sources. Table 4.6 - Analysis of alumina-bearing materials Clay

Shale

Anorthite

Nepheline

Fly ash

%

%

%

%

%

%

AI203

55-60

34

23

20-35

23-28

24-32

SiOj

3.5

45

58

45.55

45.60

45-51

2-25

2.6

6

1-3

1-3

7-11

TÍO2

1-5

2.4

CaO

0.2-0.6

MgO Na,0 + K2O H2O

0.02

10-30

13

69

Kussia has the only aluminum industry based partly on non-bauxite ruw materials, namely a nepheline syenite that contains apatite in the Kola Península. This operation was possible because of the large production of apatite for fertilizer, and the production of Portland ccmcnt as a co-product. Dilate acids are effective only in solubilizing clay if the clay is first dchydroxylated at 400°C so that the hydroxyl groups are expelled iVoin the crystalline structure in the form of water vapor, leaving bchind a porous solid having a large surface área that can be easily Icached with dilute acids. When the thermally treated silicates are atlacked by dilute acids, the metal valúes go into solution leaving bchind a skeleton of silica. For example, to render kaolinite, the main clay mineral, attacked by dilute acids, it is heated at 400°C to convert it to metakaolin: AI,(SiA)(OH),^ AIP3.2SÍO2+2Hp Metakaolin which is an amorphous aluminum silicate can be leached with acid to extract aluminum selectively leaving behind crystalline SiO^: Al203'2Si02+ 6 H " ^ 2AF^+ 2SÍ02+ SHÔ

Bauxite

Fe,03

i 'hapter 4 - Leaching Processes in Absence ofOxygen

1

<1

0.2

1

2

5-15

1-3

1-5

5

1

0.1

1-4

2-15

18

4-6

10-30

<1

trace

nil

After filtering away the silica and gangue minerals, the solution is purified, then an aluminum salt is crystallized, separated by cenirifugation, then decomposed to oxide. A variety of acids have been suggested, e.g., H2S0^, HCI, and HNO3. The problems encountered are the following: • Chemical methods or solvent extraction are used for the removal of the last traces of Fe, Ti, Cr, and V, but such methods make the processes expensive. • If HjSO^ is used for leaching and aluminum sulfate is crystallized, the cost of acid is low but the cost of decomposition of the sulfate

"IT' 70


to oxide is high. On the other hand, if HNO3 is used for leaching and aluminum nitrate is crystallized, the cost of acid is high but the cost of decomposition is low (Table 4.7). The economics of HCI leaching and aluminum chloride decomposition lies in between. • When the aluminum salt is decomposed, the acid vapors must be collected and recycled. In case of sulfate, SOj may form because of the high temperature which necessitates its transformation to SO3. In case of nitrates, nitric oxide may also form which must be collected and transformed to HNO3.

Quebec Province in Canadá and an industrial plant is planned. An advantage of the new technology is that silica is produced which can be considered as a byproduct while red mud from bauxite is a waste product. In this process, aluminum chloride is crystallized from solution by sparging with HCI gas. The crystals are separated and decomposed to AI2O3 while the mother liquor is decomposed at high temperature to recover Fe203 and HCI (Figure 4.5). HCI V

Clay

H,S04

Cost of acid

Decomposition reaction

Low

AIjíSOJj-ISHjO

Médium

HNO3

Higfi

2AIC13-6HP

Calcination

Leaching Aluminum chloride crystalization

FeÔ,

Solid/liquid separaration

Energy needed for decomposition High

HCI Calcination

Silica

^ A l A + ^SOj+ISH^G HCI

Iron chloride

u

crystalization

Table 4.7 - Main decomposition reactions during the manufacture of alumina from clays Process

71

Chapter 4 - Leaching Processes in Absence ofOxygen

-Al,03

Médium

Figure 4.5 - Simplified Orbite process

Low

LATERITES

-^ AI2O3 + 6HCI + 9Hp 2AI(N03)3-9HjO ^ A l 2 0 3 + 6HN03+15H20

Introduction Pressure leaching of clay All attempts to apply this technology cannot compete with Bayer process. The problem with Bayer process, however, is the generation of a large amount of red mud which creates disposal problem. For each ton of alumina produced one ton of red mud is generated. All this may change in the future. Researchers at the former US Bureau of Mines extracted alumina from un-calcined kaolinitic clay in 15 minutes by heating at 200°C using 20-27% HCI at 20% excess to the stoichiometric amount:

Laterites containing limonite are leached directly with sulfuric acid at high temperature and pressure. Although ferric oxide which composes the bulk of laterite is soluble in sulfuric acid, yet at high temperature ferric ions hydrolyze precipitating iron oxide and generating the acid: 2Fe3^+3H20

Fe203+6H^

Laterite in Cuba

AI2O3.2SiO2.2H2O + 6HCI -^ 2AICI3+ 2SÍO2+ SHjO

Table 4.8 shows analysis of Moa laterite. The ore is leached with H.,SO^ at 250°C and 4000 kPa in vertical autoclaves with acid-resist-

The process is being developed further by Orbite Aluminae in

ing bricks. Figures 4.7-4.8 show a view of the Moa leaching plant.

2

4

72



73

Table 4.8 - Typical analysis of latente at Moa, Cuba (dry basis)

Fe

47.5

CrOj

2.9

Ni

1.35

SÍO2

3.7

Co

0.15

MgO

1.7

Cu

0.02

Zn

0.04

AI2O3 H2O (combined)

8.5

Mn

0.8

^""'.

*

5° ^_ ,:

>-<^5T^:ifc M.l.nMS - > * V "VSlalV-

. . V'.,<>-.-' •

' ''«*1„.

JuvfHMud

Canbbean Sea

... '•

\

\

N

12.5

H;

'e ^;

LM

,,. THE BAHAMAS

a.^;í?=í^ ^ '*°° ^'^*k ^-. ^

^._-» „

ÂaiM«»»s

. -• \

Las Timas

^ ^ Z

Figure 4.8 -Acid leaching plant Pedro Sotto Alba at Moa, Cuba %

'

t=^

y NORTH ATLANTIC OCEAN '-V-M

»

A new 4-autoclave unit was added in 1998 to the already existing Ibur units in Moa, Cuba for leaching nickel and cobalt from laterites in vertical autoclaves using HjSO^.

.,

CMoimm,,

MWtor. W

Figure 4.6 - Map of Cuba where nickel is recovered

Figure 4.7 - View of the acid leaching plant Pedro Sotto Alba at Moa, Cuba (Unión de Empresas del Niquel, Havana)

Laterite in Australia In 1997 three small mining companies in Western Australia: Cawse, Bulong, and Murrin Murrin, developed processes for nickel and cobalt recovery from lateritic ores (Figure 4.9). In the three processes, acid pressure leaching technology similar to that already in operation in Moa in Cuba is used, except horizontal autoclaves instead of the vertical autoclaves are installed. Differences, however, were in the recovery step: • In Cawse process, mixed hydroxides are precipitated from the leach solution. These are then leached with ammonia, followed by solvent extraction and electrowinning • In Bulong process, mixed sulfides are precipitated from the leach solution by H^S. These are then leached in presence of oxygen, followed by solvent extraction, hydrogen reduction, and briquetting • In Murrin Murrin process (Figures 4.10-4.12), the leach solution is subjected to solvent extraction followed by electrowinning.

74


75

htpter 4 - Leaching Processes in Absence ofOxygen

Figure 4.10- Acid leaching of latentes in four horizontal titanium-lined autoclaves (courtesy Anaconda Nickel, Murrin Murrin, Western Australia) Figure 4.9 - Latente leaching plants in Western Australia

After the first five years of operation, all three projects suffered significant financial difficulties. As a result, the ownership changed and was followed also by ñame change:

Snliilíon t'rC're(t(ict¡on & NetitntliüUtio» ( A l t a 3400,' 3SIÜÍ

«Un t i l "«"'*•

• In 2001 Cawse was sold to Outokumpu Mooney Group • In 2003 Anaconda Nickel and its partner, who owned Murrin Murrin were able to raise enough funds and subsequently changed its ñame to Minara Resources • In 2004, Bulong was placed into liquidation after an apparently unsuccessful earlier attempt to raise funds. While Cawse ore preparation plant went without problems, the other two projects had problems. At Murrin Murrin, for example, the ore slurry exhibited high viscosity and had to be heated to facilítate its handling. In addition there were three ore types which needed blending for smooth operation. In the Bulong project a large proportion of the ore had to be rejected. Problems were also met with materials of construction, refractory lining of autoclaves, sulfuric acid supply, and others. These problems, however, were corrected.

1 ^ ,1

-y.

. , .

.

(An-M J300) l'ulp Plcpnruliíin lAiíii 3((W)

Prcssore Acid Uiacli AutorlavM (A)-c«320O)

-^^F*^ lltilliii{s .NeuimÜabiH

Figure 4 . 1 1 - Acid leaching of laterites in four horizontal titanium-lined autoclaves (courtesy Anaconda Nickel, Murrin Murrin, Western Australia)

76


77


I 100 200 300 mi

ttOHTHPACinC OCÍAN

afiwtor

OCtA \

Wenak'

Baigainvllkr-¿* x^ Soíomon -"!•'

.

'

^,

b

SÚIOUON

-•CS?,'.

"''í^s'rCoral Sea

Figure 4 . 1 3 - Ramu laterite plant in Madang in the Papua New Guinea iMtt Figure 4.12- Acid leaching of latentes in four horizontal titanium-lined autoclaves (courtesy Anaconda Nickel, Murrio Murrin, Western Australia)

Laterite in Papua New Guinea The $1.5 billion Ramu laterite plant is owned by Highlands Pacific and located in the Papua New Guinea 75 km south west of the provincial capital of Madang uses technology similar to Moa plant in Cuba (Figures 4.11 and 4.12). Production is 31,150 tonnes of nickel and 3,300 tonnes of cobalt annually over a 20-year mine life. The ore slurry is readily thickened to produce an autoclave feed containing 32% by weight of solids. A typical autoclave feed grade is 1.2%) nickel, 0.08% cobalt, 2.6%) Mg, and 3.4%o Al. Three autoclaves are in use operating at 250°C. An acid addition rate of 0.27 tonnes of acid per tonne of dry ore results in 96%o nickel and cobalt extraction in 60 minutes of leaching. The nickel and cobalt are then precipitated by NaOH as hydroxides. The hydroxide product is exportad for refining.

««^^1

^^^BBiÍ^.»»rfW- ** j h i t a a g ' t

p-^ ??^lü

n

i ^

A-» f" "'''i

s^.T*5VHMIiilQI^HBÍB3¿rs^

ii K

"•m - "^ ^^3^-%fc-; ' ü ^ '

L"c • v -

"\

^

'-^'"

R^

'^^î, ~'Sl ^ ^

"" |hi¿¿>î¿^ "^ -'^^' i \

j

"''-^MÍ

íjp::f••' <- '-.

é'

W' '^

p¥ "••'M-

Figure 4.14 - Ramu plant

The upgraded ore is pumped, as a slurry, through a 134 km pipeline to a high pressure acid leach processing facility at Basamuk Bay on the Rai Coast. A sulfuric acid plant producing 3,350 tonnes per day of 98.5%) acid supplies the acid requirements. Pilot tests were conducted at Hazen Research in USA and Lakefield Research Laboratories in Canadá.

78


Laterite in Madagascar

—I I

Another laterite deposit is being exploited in Madagascar using high pressure acid leaching technology similar to Moa plant in Cuba. At ^ • Toamasirtc the mine site, an ore preparation Antananarivo < Ambatovy plant, a pressure acid leach plant and refinery, and port facilities on the coast were constructed. • Fort-Dauphir The ore is transferred to the coast through a 220 km slurry pipeline. Figure 4 . 1 5 - Location map of The project known as Ambatovy Ambatovy mine in Madagascar (Figures 4.15-4.17) is owned by Dynatec, Sumitomo, and SNC Lavalin. Annual design capacity is 60,000 tonnes of high grade nickel, 5,600 tonnes of high grade cobalt, and 210 000 tonnes ammonium sulfate for at least 27 years. Project investment is estimated at $ 4.5 billion.

79


««í^-

•¡.HMUQIÍOR

murmujSKnoH

couwnMcmunr ««MMWIOW

SULfHOe

JS*_

::,...:..J-

NICKtL

Figure 4.17 - Production of nickel, cobalt, and ammonium sulfate fertilizer from laterites in Madagascar

Figure 4.16- Autoclaves at Ambatovy

The pressure acid leach of laterite ore takes place in 5 parallel units each consisting of a feed tank, slurry heaters, two Geho PD pumps, an autoclave and flash vessels. The ore slurry received from the thickeners is heated with steam in a series of direct heaters and pumped, continuously, at a temperature of up to 200°C into the autoclaves where it reacts with sulfuric acid and oxygen at 260°C. The operation is similar to Moa plant in Cuba with the exception that the

80


autoclaves are horizontal instead of vertical. Three autoclave are 40 m long and 7 m diameter, divided into seven compartments, made of titanium ciad steel and weighs 700 tonnes while the other two are smaller, composed of only four compartments. Three autoclaves were manufactured in China and two in Belgium.

ILMENITE Introduction i'he major titanium mineral is ilmenite, FeTiOg. In the early method «f TÍO2 pigment manufacture the ore was treated with concentrated HjSO^ at 110-120°C to form ferrous sulfate and titanyl sulfate:

Hydrogen sulfide is used to precipítate nickel and cobalt from the leach solution. The precipítate is then leached by ammonia using Sherritt-Gordon technology of Fort Saskatchewan in Canadá to get a solution from which nickel and cobalt are separated by solvent extraction. Puré nickel and puré cobalt are then recovered by precipitation with hydrogen under pressure and ammonium sulfate crystallized from the remaining solution to be marketed as fertilizer.

FeTiO +4H"

Fe2^+Ti02^+2H20

After removing the insoluble residue by filtration, the solution containing 120-130 g/L TÍO2 and 250-300 g/L FeSO^ was concentrated under vacuum at 10°C to crystallize FeSO^*7H20 which was then centrifuged. Titanium oxide is then precipitated from solution by dilution and seeding resulting in the formation of dilute H2S0^ for disposal (Figure 4.19).

Effect ofCl ion Saline water was found to enhance the dissolution of nickel from New Caledonia laterites. The use of sea water resulted in a chloride concentration of 7.8g/L in the autoclave feed. Leaching tests were conducted at 30% solids, temperature range 230-270°C. These results, however, did not apply to Western Australia laterite ore.

Conc. H.SO,

I H,0

Ilmenite

Baking

i

Laterite in Turkey

Leaching Filtration

Pressure leaching is used in Manisa Gordes near Ezmir in Turkey by Meta Nickel Cobalt Company. It is the first autoclave installation in Turkey. The autoclave was constructed in China, is 32.5m long, 7 m diameter and weighs approximately 580 tonnes (Figure 4.18). The laterite ore containing 1.26% Ni and 0.08% Co is treated with H2SO4 at high temperature (250°C) and pressure (40 bar) in the autoclave. Initial investment is about 3 billion dollars figure 4.18

Autoclave in Turkey

81

('hapter 4 - Leaching Processes in Absence ofOxygen

Residue

Crystallization Centrifuge

ni

Seed

- FeSO^ - 4H2O H,0

Hydrolysis Filtration

Dilute H,SO^ Z

4

Drying

i

Calcination TÍO. /I

1 n

'T'i

i_.

ii

j i?_

r.

•T:/~\

82

Pressure Hydrometallursv

l
Methods of treating ílmenite Because of the pollution problems associated with the disposal oi" FeSO^ and dilute sulfuric acid, iron in the ore was separated at an early stage by a pyrometallurgical method developed in 1950s. The ore was mixed with just enough anthracite to reduce the iron oxide component of the ore then charged in an electric furnace where iron oxide was reduced to metal while titanium was separated as a slag:

83

ilment process of the Sorelslag, however, still suffered from the ¡H)sal problem of the waste acid. As a result two processes were cloped: 111 l'roduction of synthetic rutile. This technology was introduced in Ihc 1960s which involved leaching of ilmenite in autoclaves by 20% HCI at 120°C and 200 kPa and obtaining a residue rich in titanium (90 95%) TÍO2) known as "synthetic rutile" (Figure 4.21):

FeTiOg + C -> Fe + CO + TiO

FeTi03+2H"

2(slag)

T¡0_[¡mpure] + F e 2 ^ + K 0

Fe203+3C-^2Fe + 3CO HCI

The slag (Table4.9) is mainly iron magnesium titanate, {Fe,MQ)T\^0^^, and a small amount of silicates. In Sorel, Quebec it is called Sorelslag. It is high in titanium and therefore preferable to ilmenite in manufacturing pigment or metal.

Ilmenite

.. i

Digestión V Filtration

Synthetic rutile

" Table 4.9 - Analysis of titanium raw materials Rutile

Ilmenite

%

%

TiOj

80-95

TÍ2O3

0

Sorelslag

%

Synthetic rutile, %

43-59

72.1

90-95

0

FeO

10-20

0

9-38

Fe,03

8.9

5-25

0.0

0.0

0.2

Fe SiO,

0.4-4.0

5.8

AI2O3

1.3-3.3

6.5

MgO + CaO V

0.1-4

7.3

0.4-2.0

0.4

.

Although the electric furnace treatment of ilmenite eliminated the bulk of iron, the slag produced in Sorel was only about 72% T\0^. It was only suitable for treatment by sulfuric acid to produce pigment. It was not economical to be treated by chlorine to produce pigment because it still contained much impurities and the process still suffered from the disposal problem of the waste acid. The sulfuric acid

I

Oxyhyclrolysis

Figure 4.20 - Production of synthetic rutile from ilmenite

'í"he synthetic rutile is then treated by chlorine to prepare TICI^ from vvhich TiOj or titanium metal are obtained while ferrous chloride is treated by oxyhydrolysis to obtain iron oxide and HCI for recycle: 2FeCl2+2Hp + y202-

Fe203 + 4HCI

12] Upgrading of Sorelslag. QIT Per et Titán at Sorel installed in 1980's a pressure leaching plant to upgrade the slag to 95%) TÍO2 by heating with HCI at 150°C. This treatment removed MgO, CaO, and Fe203 but did not remove silica which still remained in the slag.

1

84


!>ler 4 - Leaching Processes in Absence ofOxygen

WOLFRAMITE AND SCHEELITE Introduction

85

FeWO^+ 2 0 H - ^ W 0 / - + Fe(0H)2 Scheelite is decomposed by sodium carbonate solution at 225°C:

Wolframite, (Fe,Mn)W04 or (Fe,Mn)O.W03, and scheelite, CaWO^ or CaO'WOg, are the most important sources of tungsten. Table 4.10 shows a typical analysis of concéntrales containing these minerals. Both materials can be decomposed by either acids or alkaline Solutions. Table 4 . 1 0 - Average analysis of tungsten concentrates Wolframite

CaW04+C032-

I he formation of CaCOg films on the mineral particles retards the caction. This factor can be eliminated, however, by using rotating iilociaves containing steel balls. The solution of sodium wolframate •. purified by precipitation with acid: WO/-+2H^^W03-H20

Scheelite

WO3

75-65

70-78

FeO

5-15

0.4-2.0

MnO

5-20

Cao

0.2

0.1-0.2

17-19

W0/-+CaC03

VACUUN RLIER

ÜtrCfi

l^s.

Acid digestión is usual ly conducted with concentrated HCI in excess: FeW04+2H^

Fe2-+W03-H20

CaW04 + 2H"

Ca2^+W03-H20

Sulfuric acid cannot be used because of the formation of insoluble calcium sulfate. The digested mass is washed with water to remove iron and manganese chlorides the residue is then dissolved in hot NHÔH. Ammonium wolframate is crystallized from the solution by evaporation. Alkaline leaching Leaching of wolframite with concentrated NaOH is conducted at high temperature in an autoclave to yield a solution of sodium wolframate, while iron and manganese are precipitated as hydroxides:

TJl

-• SOUTIOH OF '.OOIUN TüHCSI*TE

Figure 4.21 - Tungsten concéntrate processing in autoclave

PYROCHLORE Introduction l'yrochlore, (Ca.BaP'Nb^Qg'NaF, (Table 4.11) is mainly used to prepare ferroniobium by pyrometallurgical method. To prepare metallic niobium a puré oxide is prepared first by treating the concéntrate by hydrometallurgical method.

86


Table 4.11 - Typical analyses of niobium concentrates Pyrochlore [%] Quebec

Brazil 60

NbjOj

68.7

TaPs

0.2

FeO

0.4

4

_

MnO CaO

14.8

10.2

BaO

—

16

MgO

0.5

i/>tcr 4 - Leaching Processes in Absence ofOxygen

87

ARSENIDES AND ANTIMODES liilroduction I l)c presence of arsenic and antimony in copper sulfide concentrates IS uiidesirable because these metáis complícate the smelting and rclining of copper. As a result there is interest to remove them beforc smelting. One route is leaching the concéntrate by an alkaline «odium sulfide solution at high temperature and pressure.

SnOj TÍO2

0.6

I'urification of chalcopyrite concéntrate

WO3 Rare ea rths

2.0

F

3.9

Na20 + K2O

7.3

Treatment of pyrochlore Pyrochlore can also be beneficiated to a product containing 90-97% NbPg by reaction with 36% HCI at 200°C and about 1000 kPa for 4 hours in a pressure reactor. The reaction is basad on the formation of the niobium ion which hydrolyses to Nb205 at the reaction temperature. The reaction takes place in two consecutive steps:

A copper sulfide concéntrate containing 4% As and 7% Sb was Ircated in British Columbia by Equity Silver Company by this mcthod. The finely divided concéntrate is leached for 16 hours at ! 10°C to solubilize arsenic and antimony sulfides: AS2S3+3S2-

2ASS33-

Sb^Sg+SS^-

2SbS3^

After filtration, the copper concéntrate is shipped to smelters. The leach solution contains 30 g/L As and 53 g/L Sb. It can be treated in two ways:

SÍNbPs'CaO) + 2HC1 -> 2Nbp^ + Ca^NbÔ^ + CaCI^ + Hp pyrochlore

• Electrolyzed in a diaphragm cell to get antimony and regenérate the leach solution:

C a ^ N b p , + 4HC! -> Nbp^ + 2CaCl2 + 2H2O

SbS32-+4e-^Sb + 3S2-

Calcium niobate, Ca2Nb20-„ is formed as a non-porous intermedíate product on the pyrochlore grains through which the reactant and the products must diffuse.

• Treated with oxygen in autoclaves at 150°C and 550 kPa to decompose the antimony thiocomplex: Na3SbS3+ 4NaOH + H_0 + '^lO^-t NaSb(OH).+ SNa^SO,

n" 88


In the flash evaporator, precipitation of sodium antimonate, NaSb(OH)g, takes place; it is filtered off and recovered. Arsenic remaining in solution is then precipitatedby lime in another autoclave at 1600 kPa oxygen pressure to precipítate calcium arsenate:

Leaching Processes in Presence of Oxygen

2Na3AsS3+ 3Ca(OH)2+ 130^ -^ Ca3(AsOJ,+ 3Na2SO,+ 3H2SO,

This is filtered off and packed for disposal. The remaining solution containing sodium sulfate is evaporated to crystallize NajSO^'IOHjO. Due to the presence of traces of arsenic in the crystals, these are re-dissolved and retreated with lime in autoclave to precipítate the remaining arsenic. Puré sodium sulfate is then obtained by crystallization. The plant, however, was shut down for economic reasons. In a similar way complex cassiterite, SnO^, concéntrate especially those from Bolivia was purified by boiling at 110°C with HCI in autoclaves to remove impurities. This was conducted in rotating spherical autoclaves at the Longhorn Smelter in Texas. This resulted in removing most of the impurities and the tin oxide obtained was amenable to conventional smelting.

Uranium Oxides 89 Introduction 89 Pitchblende, Carnotite 90 Leaching ofUO^ 90 Sulfides 97 Introduction 97 Leaching in ammoniacal médium 98 Leaching in neutral and acid médium 102 Liberating of nickel and cobalt from pyrrhotite and arsenopyrite. 119 Liberating of gold from pyrite and arsenopyrite 124 Selenides and Tellurides From Anodic Slimes 135 Introduction 135 Acidprocess 136 Arsenides 140 Unsuccessful Pressure Leaching Processes 147 Clearprocess 147 Sherritt-Cominco process 152 Lurgi-Mitterberg process 152

URANIUM OXIDES Introduction Uranium occurs in nature mainly in the form of an oxide. Although it forms numerous oxides (Table 5.1), only two are the most important: UO2 and U30g because they constitute the bulk of uranium

90

Pressure Hydrometallm

TSÁ

ores. Uranium trioxide, UO3, is soluble without the need of an oxidizing agent. Uranium in this oxide is in the hexavalent state; ii does not however occur in nature in the free state, but in association with vanadium and potassium in the mineral carnotite, K2O.2UO3. V20g. Carnotite is readily soluble in acids in the absence of oxidizing agents. Table 5.1 - Uranium oxides Oxide

Valency

UO,

uraninite

U.Os U3O3

Natural form

4,6

UO,

/,•/•

.^

U30g+ 2 H ^ ^ U2O5+ U022^+ H2O

U2O5 + 2H* -. UO2 + UO^^ + Hp I he formula U02»2U03 should not be taken as indicating the prest-neo of two types of uranium in U30g. X-ray analysis shows that all uranium atoms in UgOg occupy equivalent positions; there is probgbly a resonance between two (or more) valency states 4 and 6.

Solubility in dilute H^ SO^ in absence of oxygen insoluble

does not occur in nature

partially soluble

pitchblende

partially soluble

occurs only in combination with vanadium oxide as the mineral carnotite

soluble

100

1

—\

1

1

1 (B)

80

c o '% 60

(A)

X Lll

E 40 g 'c

Leaching of UO^ Uraninite, UO2, is insoluble in dilute H2S0^, and uranium in this oxide occurs in the tetravalent state. Pitchblend, UgOg, is partially soluble in dilute H2SO4; uranium in this oxide occurs in both the hexavalent and tetravalent states and may be represented as UO2'21103. ^^^^ accounts for the fact when dissolved in dilute H.SO^ in absence of 2 4 air, mixtures of uranium(IV) and uranium(VI) are obtained: U303 + 4 H ^ .

91

• Leaching Processes in Presence of Oxygen

20 i ..

1

8

1

1

12 16 Time, Hours

1

1

20

24

Figure 5.1 - Solubility of a pitchblende ore sample containing 0.22% UjOg in dilute HjSO^. (A) In absence, and (B) in presence of oxidizing agents. Plotted from data in a Canadian report (Anonymous, 1955)

U02+2U022^+2H20

The reaction, however, seems to be more complex because the composition UO2 is never reached; in practice a máximum dissolution of about 58% is reached as shown in Figure 5.1 and not 66.67% as expected according to the above equation. It seems that the intermedíate oxide U2O5 is formed and the product is a mixture of UO2 and U2O5:

As mined, pitchblende contains about 1% U3O3, but it can be easily concentrated by gravity methods to 50% UgOg. The main occurrences are in Joachimsthal (Czechoslovakia), Shinkolobwe (Zaire), Elliot Lake (Ontario), and Athabasca Lake (N.W. Canadá). Thucholite is a uranium mineral containing thorium, carbón, hydrogen, and oxygen that occurs mainly in South African gold ores. These ores average 0.02-0.1%) U30g and are processed first for the recov-

92


ery of gold. Davidite is another uranium mineral containing iron, cerium, titanium, vanadium, chromium, and zirconium that occurs mainly at Broken Hill, Australia. It is a refractory mineral difficult to dissolve. At Palabora in South África, uranium is associated with copper sulfides; it is recovered from the flotation tailings by gravity methods as a concéntrate containing 2.5-5% UgOg mainly as the mineral urano-thorianite. Uranium recovery as a by-product of copper oxide leaching operations has already been referred to on page Leaching agents commonly used are the following. In Australia, a uranium deposit containing 0.06% UgOg and 2.1% Cu as sulfide is under exploitation at Olympic Dam. A copper sulfide concéntrate is obtained by flotation leaving a residue containing the bulk of uranium and 0.3% Cu. Both copper and uranium are leached from the residue by acid in presence of Fe^^ ion. On the other hand Northern Saskatchewan in Canadá became the world center for uranium industry: large deposits are under exploitation at Key Lake (1.5% U3O3), Cigar Lake (13.6% UgOg), and McArther River (18.7% U30g). Sulfuric acid, either dilute for easily soluble uranium minerals, or concentrated for the refractory minerals, is the most commonly used acid. Leaching may be represented by the following equations:

V20^ + 2W+2e--^Hp Overall reaction:

\^0^ + 2W+'A0^-

UO/^+Hp

Negatively charged sulfate complexes are formed, e.g., [\JOJ^SO^^. Oxygen or other oxidizing agents such as Mn02, NaClOj, or NaNOg are commonly used. Uranium is recovered from solution by ion

('hapter 5 - Leaching Processes in Presence of Oxygen

93

cxchange or solvent extraction. Alkali carbonate process Alkali carbonate process is used when the ore contains appreciable amounts of acid-consuming gangue and it is conducted at high tempcrature and pressure in autoclaves. The reactions that take place in this case are: U02ÛO/^+2eUO 2-+ 30032-^ [U02(C03)3} y202+H20 + 2 e - ^ 2 0 H -

Overall reaction: UO2+ 3CO32-+ V\p + 'ÁO^-^ [U02(C03)3]^+ OH-

Since OH" ion is formed during leaching and there is a possibility that insoluble uranates may be formed, sodium bicarbonate is usually added to the solutions to prevent such side reactions: HCO-+OH-

CO32-+H2O

Mgure 5.2 shows pressure leaching of uranium ores with sodium carbonate at Beaverlodge, Canadá. Ammonium carbonate leaching under pressure has the advantage of having less attack on silicate minerals and on alumina and uranium can be precipitated by stripping with steam to decompose the uranium complex; the evolved NH3 and CO2 are absorbed and recycled. It is used together with hydrogen peroxide for in situ leaching of underground uranium ores, e.g., in Texas.

94 Pressure Hydrometallursv ^ ^ ^ ('hapter 5 - Leaching Processes in Presence ofOxygen

95

TO ímoSPHERE lEWHfKBPuiP

SPUSB TOKÜ

Figure 5.4 -Atlas Minerals

fllTES SWE AUroCUVES

' mi

Figure 5.2 - Pressure leaehing of uranium ores with sodium carbonate at Beaverlodge, Canadá

operating at 330 kPa and 120°C for 6 hours.

LasVegas Figure 5.3 - Location of Moab on Colorado River

•

autoclaves

RIS0 National Laboratory in Denmark used tube autoclaves (Figure 5.5) at 290°C for leaching uranium ores from Greenland. The ore contains fluorides which are then removed for disposal as calcium fluoride by adding gypsum to the solution: CaS0^.2Hp + 2F-

CaF2+SO/-+2H20

Leach residues. Residues from uranium extraction plants using either HjSO^ or Na2C03 contain all the radium originally present in the ore. These residues are at present stock piled because radium is not in demand. Radium decays into the radioactive gas radon. The diffusion of this gas in the environment, the scattering of radioactive dust particles by wind, and erosión of the piles of residues by water, represent a serious pollution problem. A typical disposal pond contains 0.6 mg radium per ton of solids. Abandoned plant sites are particularly hazardous because tailings dams may either erode or rupture and reléase tailings to streams. Therefore, controlled storage of uranium mili residues must be maintained after the life of the plant to safeguard the environment from radioactive pollution. Considering the 1622-yearhalf life of radium 226, storage must be controlled for many thousands of years to enable abatement of the radiation hazard by natural decay of radium and its products. For this reason, leaching uranium ores with HNO3

96 PressureHvdromP.tnlh.rcr.,

or HCI followed by precipitation and separation of (Ra,Ba)SO^ by adding BaCl2 is being considered as a mean to solve this problem although these acids are more expensive than the commonly used

. hapter 5 - Leaching Processes in Préseme ofOxysen

97

A more practica! solution, however, is to fill the open pit mine with water and deposit the tailings and residues at the bottom, thus the water above will act as a protective layer against radiation (Fig ure 5.6).

• igure 5.6 - Tailings and residues from uranium treatment plant being stockpiled under water in an unused open pit mine

SULFIDES Introduction l.eaching of sulfides in presence of oxidizing agents may lead to the formation of sulfates or elemental sulfur. In neutral médium leaching is slow at ambient conditions, but rapid at high temperature: MS^M2"+S22

... ..

xmmm^'J

Figure 5.5 - Tube autoclaves for leaching uranium ores

4

()veralIreaction:MS + 2 0 „ , ^ M S O ^ 2(aq)

4

in acid médium and at temperature not more than 150°C elemental

98

i 'hapter 5 - Leaching Processes in Presence ofOxygen


Copper, nickel, and cobalt form soluble ammine complexes with ummonia. The process has mínimum corrosión problems and any pyrite present will not be attacked. In this process all the sulfur is oxidized and recovered as ammonium sulfate and marketed as fertihzer. The overall reaction is:

sulfur is formed: M S - * M 2 ^ + S + 2e-

V20^+2W+2e'^Hf>

MS + nNH3+202-

Overall reaction: MS + 'ÁO^+IW ^W^+S + V\p

2 H " + S 0 2-

[M(NH3)J2^+SO,

where M = Cu, Ni, or Co. It is important to control the amount of free ammonia in solution otherwise higher ammines like cobalt hcxammine complex, which is insoluble, will be formed. Analysis of the concéntrate treated is given in Table 5.2.

To avoid the deposition of liquid sulfur on the sulfide and thus retarding the reaction, a small amount of coal or a surface active agent like lignosulfonate or Quebracho is added. Above 150°C elemental sulfur oxidizes to sulfate:

S + VAO^+Hp

99

I

Table 5.2 - Pressure leaching of Sherritt-Gordon sulfide concéntrate

4

In basic médium oxidation of sulfides takes place in several stages and any of the intermedíate compounds: polysulfide, thiosulfate, etc., may be present in the leach solution. The leach solution is also free from iron, since ferric oxide is precipitated.

Ni Cu Co Fe S Insol.

Leaching of nickel and liberation of gold from sulfide minerals have received the greatest attention using pressure leaching. For example: leaching of nickel from pentlandite, (Fe,Ni)gSg, and its liberation from pyrrhotite, FeS, and liberation of gold from pyrite and arsenopyrite.

Residue, %

10-14 1-2 0.3-0.4 33^0 28-34 8-14

0.6-1.4 0.2-0.3 0.1-0.2 42-52 9-15 12-16

The process involves the steps shown in Figure 5.7. Leaching The concéntrate is mixed with water and ammonia and leached in autoclaves under air pressure of 700 kPa and at 70-80°C for 20-24 hours. Reaction is exothermic and therefore extra heating of the autoclaves is not required.

Leaching in ammoniacal médium Sherritt-Gordon process is the first process on industrial scale that uses autoclaves for leaching sulfide minerals. The temperature is only 80°C but pressure is used to increase the solubility of oxygen in solution henee increasing the rate of leaching. The method has been used successfully since 1953 for the treatment of Ni-Cu-Co sulfide concéntrate on large scale at the Sherritt-Gordon Plant in Fort Saskatchewan, Canadá.

Fiotation concéntrate, %

M

¡'urification The leach solution contains beside nickel and cobalt ammines, excess ammonia, copper, thiosulfates, and thionates. Ammonia is removed by distillation and is recovered in scrubbers. During distillation

Pr.essure Hydrometallurgy

100

W&'hapter 5 - Leaching Processes in Presence ofOxygen

101

most of the dissolved copper is precipitated as sulfide: S3O/-+ [Cu(NH3)J2-+ 2H2O ^ 2 S 0 / - + CuS + 4 N H / S2O32- + [Cu(NH3)4r+ Hp -^ S 0 / - + CuS + 2 N H / + 2NH3

After filtration, the residual copper (about 1 g/L) is precipitated by a controlled amount of H2S in autoclaves at 130°C. This second precipítate contains some NiS and is recycled to the leaching stage. T3 U

s

Ni-Co-Cu sulfide Concéntrate

13 o

h*

Air

NH3-

S 'S o

Leaching í 80 C. 700kPa

Filtration ^ Y

i \ • •" *" Residue: gangue \ Fe (0H)3, PbS04, precious nnetals

o o

Boiling o t3 Filtration

-•- CuS

HaS —

o

o

íPrecipilation t. of traces of Cu^*

Filtration Air

CuS, NiS, CoS recycle to leaching Circuit

-O

u Q

Oxidation

00 Filtration

»-Fe(OH)3

;-< Purified ammoniacal ammonium sulfate solution containing 45g/LNiand1g/LCo

Figure 5.7 - Pressure leaching of Ni-Co-Cu sulfide concéntrate; the Sherritt-Gordon process

Oxyhydrolysis In this step oxidation of thionates and hydrolyzing sulfamate takes place. The presence of thiosulfates and thionates in a nickel or a cobalt solution is undesirable because it leads to contamination of

3

102


the fertiHzer produced later. For this reason the copper-free solution is then digested at 175-200°C in an autoclave in the presence ol compressed air at 4200 kPa for two reasons: To oxidize thiosulfates and thionates to sulfates: S2O32-+ 2O2+ 2 0 H - ^ 2SO/-+ H p

1

iper

rrilt-Gordon chalcopyrite process. Pressure leaching technology becn developed in 1960s for sulfide concentrates essentially by rritt-Gordon Mines in Canadá [now known as Dynatec] operat..^ m a sulfate system (Figure 5.9): CuFeS^+r/zO^+aH^-

• To oxidize traces of ferrous ion to ferric which is hydrolyzed and precipitated. H.SO,

The purified solution upon clarification contains 45 g/L nickel and Ig/L cobalt as ammines, and ammonium sulfate. Recovery , This involves the precipitation of metallic nickel by hydrogen, oxH dation of Co^"^ to Co-^* by air, then precipitation of metallic cobalt bj hydrogen. The remaining solution is evaporated and the crystals o ammonium sulfate separated and sold as fertiHzer. Precious metalj if present, remain in the residue and may be recovered by a separat leaching cycle.

103

• ipUr 5 - Leaching Processes in Presence ofOxygen

Cu2^+FeOOH + 2S + H20

Cop per ;once ntrate — Oxygen V

^

r \

AQ Oxldatlon

" Filtratlon

Residu e

"

'

Electrowlnnlng

Flotatlon

Copper

Cyanldatlon

1

Elemental sulfer ^

Residuo to waste

Precious metáis

Leaching in neutral and acid médium Figure 5.9 - Pressure leaching of copper sulfide concentrates

There a number of processes that uses autoclaves for leaching sul fides in neutral and in acid médium (Table 5.3). As mentioned earlief below 150°C elemental sulfur is formed while above that tempera' ture sulfate. The case of MoSj is special in that molybdenum ion hydrolyses during leaching forming molybdic acid. Table 5.3 - Leaching of sulfídes in neutral and in acid médium Metal recovered Copper Zinc Nickel, cobalt Molybdenum Platinum metáis

Process Sherritt process, Freeport McMoran [formerly Pheips Dodge], Anglo American, Chínese process, Halex, Sepon, Telfer Sherritt process Nickel matte, nickel-cobait sulfide precipítales Transforming M0S2 Into molybdic acid Platsol process

process is conducted at about 150°C and at 1500 kPa. Coal Idcd during leaching in the amount of 20 kg / tonne to prevent lenlal sulfur from adhering to the chalcopyrite and retarding the hing. This process has the advantage that iron is separated as tisoluble residue because of the oxidation of ferrous ion to ferric its hydrolysis: 2Fe2^ + 2H" + 720^ -^ 2?e^' + Hp Fe3^+ 2H2O -^ FeOOH + 3H"

104


( hapíer 5 - Leaching Processes in Presence ofOxygen

The residue from the leaching operation, after flotation of sulfur, should be agglomerated with Portland cement and stockpiled on an impervious base, in the form of dumps to be treated by cyanidation for precious metáis recovery. The process has the foUowing characteristics: ^ • The oxidizing agent does not need regeneration. • The iron component of chalcopyrite is obtained as a residue during leaching. • Selenium and tellurium will be associated with the elemental sulfur while arsenic precipitates as ferric arsenate • The precious metáis in the concéntrate could be recovered from the residue. • The process is self-sufficient with respect to the acid used when the copper-containing solution is electrolyzed. In the first step, leaching is conducted under mild acid conditions (pH about 3) and in presence of a mixture of HCI and H2SO4. Under such conditions copper hydroxysulfate is formed which is solubilized in a second step at atmospheric presure in dilute H2SO4. There is no obvious advantage, however, in this process as compared to Dynatec. In both processes, the precious metáis are recovered from the residue by cyanidation after the flotation of elemental sulfur. Freeport McMoRan chalcopyrite process. In 2004 Phelps Dodge [now Freeport McMoRan] built a plant for pressure leaching of chalcopyrite concéntrate at 200°C to get copper in solution and genérate sulfuric acid for heap leaching - solvent extraction of oxide operations at Bagdad, Arizona (Figures 5.10 and 5.11). Few years later another plant operating at 150°C was constructed to recover elemental sulfur (Figures 5.12 and 5.13). The process is not different from the Sherritt process described above. The location of the two mines is shown in Figure 5.15.

Figure 5.10 - Pressure oxidation of chalcopyrite at Bagdad, Arizona

Concéntrate Sluri

Figure 5 . 1 1 - Pressure oxidation of chalcopyrite at Bagdad, Arizona

105

106

Pressure Hydrometallurg¡^

Copper sulfide concéntrate

•r

Water 1'

)/)><•'/• 5 - Leaching Processes in Presence ofOxygen

107

Sulfide concéntrate H,SO,

"-'2

±1^°'

Leaching, 220°C

Leaching

V

.

S/L Separation y

\' Solvent Extraction

Organic

r

1

+

Filtration

solution Purifi catión Barren solution

Heap leaching

''

^r

I

Electrolysis

Metal

Stripping

1

Gaugue, S, FeOOH, PbSO^

Acid

l'igure 5.14 - Recovery of copper from sulfide concéntrate by leaching under pressure with elemental sulñir recovery

T Copper

Figure 5.12 - Recovery of copper from sulfide concéntrate by leaching in autoclave at high temperature

Figure 5.13 - Pressure oxidation of chalcopyrite at Morenci, Arizona

Figure 5.15 - Major copper mines in Arizona showing Bagdad and Morenci [X]

108


Anglo American processes Anglo American Corporation of South África and the University of British Columbia in Vancouver, Canadá has developed a mediumtemperature pressure leaching process for the extraction of copper from chalcopyrite concéntrales. Leaching is performed at 150°C and 700kPa oxygen pressure. A key element in the process is to get máximum yield of elemental sulfur (about 60%) through the use of the surfactants lignosulfonate and Quebracho. A combination of these two has been found to significantly promote the kinetics of copper extraction.

Chapter 5 - Leaching Processes in Presence ofOxygen

was recovered in the elemental form, and only traces of arsenic were present in solution. Copper was recovered from the leach solution by solvent extraction using BK-992, an organic solvent produced in China, followed by electrowinning. After the pilot tests a commercial plant is planned. Copper concéntrate

, 1 , ,. i V

Oxygen

CESL chalcopyrite process Developed by Comineo Engineering Services Limited in Trail, Canadá will be used in Brazil by Compania Vale do Rio Doce known as CVRD. The process is in two leaching steps: the first step in dilute sulfuric acid containing NaCI and oxygen at 150°C. The conditions are so chosen to be less corrosive; thus at pH 3 chalcopyrite is decomposed forming copper sulfate hydroxide, ferric oxide, and elemental sulfur. After solid-liquid separation, the solids are subjected to a second stage leaching at atmospheric pressure and pH 1.8 to selectively dissolve the copper sulfate hydroxide. Copper sulfate solution obtained is then extracted by organic solvents and the strip solution electrolyzed to produce copper cathodes (Figure 5.16). Chínese chalcopyrite process Hydrometallurgical process for complex chalcopyrite at low pressure and temperatura by workers at the Beijing General Research Institute of Mining & Metallurgy and the University of Science & Technology in Beijing, China. It was found that at 110°C, P^^ = 500 kPa, 30 g/L H2SO4 and 30 g/L [01"] ion, copper extraction was more than 95%, little pyrite was attacked, more than 90% of the sulfur

Make-up acids (HCI + H,SO,)

Pressure leach 'f

Filtration

A pilot plant started in 2003 to recover copper from chalcopyritechalcocite concéntrate known as Mantos Blancos concéntrate by Anglo American-Chile. Leaching was conducted at 180-220°C is presence of oxygen and promising results were obtained.

109

-^

' J V Solids Atmos. leach "

Raffinate

Filtration y Solution Solvent extract •M-

Loaded

Organic ,^

Flotation

" Separation

^ Residue (FeÔj + Gangue) to waste —•^ Elennental sulfur

1 Precious metáis

Stripping ip Solution

if

L+-

Electrowinning Pure copper

Figure 5.16 - CESL process

Halex process In an attempt to solubilize simultaneously copper and any gold present in chalcopyrite, the Halex process was developed by Intec Copper Proprietary in Australia. The leach solution contains large amounts of NaCI and NaBr in circulation between the copper electrowinning stage and a diaphragm cell in which the following reaction takes place at the anode compartment:

Br-+2C|-^BrC|-+2e-

sssoojd uodag JO isaqsMoy p^liBi^Q - g r g sinSij

^ « j soiwnivi

T=^PHMII|

.-G^BS-,

DaM

oj

s¡

r^

^=i o t/i

c c3 r.

x: 13

-4->

+

•rf

O

:3

o

1 L—

c

+ (n rf

1

0)

co

+

kl

CO

C3

+

^ O

CQ

O J

++

<

<

c

c o

CM

O — O1Q u

T

c/3

T 1 CM

O

u '^ +

Z5

<

Q. Q.

ctí

W

O

c c c o 0)

(/) (U t-

•'— (i>' o c

+ <

> ^ >. PH 1

(D

3

^

bX) 00

o

o

S &. PH

o

o

00 -tí

c

l-H

O

03 (U
(/2

••->.

_o

o

(U "O

o

bX) o C o

os

:3 c 03

W

X

c

-T:;

3

O., 1

0)

'

ro 0

O

c o

00

(0

I - ^

1!

-a

cu

o

c

c

o ü)

ro

ü

w

S

s

o ü

'

c

'S c -o Q

g

00

"¿

M

-(—» c o o

C/D

•>-•

13

3 o %-, "o

o

c o

_g

o

1

O

C!) C

M

*j

O)

-T3

+0 LL (M

¡3

[i.

•;-

-O

1

(N

00 ID

X

o

05

r~o ^/^

O o

-3¿

-a

en

o

X

O (U

s ctía c^ ^ -

> o

o

'

(U (U ao -O o oj u

^ (1>

o

^ M

-o c o -o o c o "o o

IS

(U en

o
(D

(O

TJ

t-

03

.ti ti (U

>



113

Make up H2SO4

•essure leaching stage is conducted in autoclave at 180-195°C er oxygen pressure.

Sulfide concéntrate i

_£

olid-liquid separation the pyrite residue is stockpiled and ition is sent to the atmospheric leaching stage. A detailed et is shown in Figure 5.18.

Leaching

1.

Filtration

Oz

• Gangue, S, PbSO^ FeOOH

Purification

'oject

\il studies were conducted to treat the gold-chalcopyrite ate known as Telfer project in Western Australia using leaching technology. The sulfide feed was separated into ; one rich in chalcocite treated at 100°C and the other rich oyrite treated at 220°C. The leach solutions then combined essed in a solvent extraction - electrowinning plant to Dpper while the residue is treated in a cyanidation - acti-coal - electrowinning circuit. The decisión not to proceed th a commercial plant was based on the fact that an underDosit was discovered and the sale of the concéntrate would ;onomical.

Spent electroiyte

Electrolysis

Y Metal

Figure 5.19 - Flowsheet for the aqueous oxidation of sulfide concéntrales in acid médium

tn of zinc sulfide concéntrate is conducted at 150°C and ygen pressure: ZnS + 2W+ YzO^^ Zn2*+ S + Up > is now used Cominco's refinery at Trail, British Coida in 1981 (now the largest zinc produces in the world), solves two problems facing the hydrometallurgical zinc nc goes into solution because no ferrites are formed. ess is independent of fertilizer manufacture because no ide is formed.

Figure 5.20 - Pressure leaching plant of zinc sulfide concentrates at Comineo, Trail, British Columbia (Sherritt-Gordon)

114


Five plants were installed later after the one at Trail: at Kidd Creek División of Falconbridge in Timmins, Ontario, at Ruhr-Zink in Datteln, Germany, at Hudson Bay Mining & Smelting in Flin Flon, Manitoba in 1993, and at Kazakhmys Corporation at Balkhash in Kazakhstan in 2003. The last one was based on Sherritt's two-stage pressure leaching technology (Figure 5.21). The two-stage process operates also at 150°C but results in high zinc extraction, a solution with low acidity suitable for electrowinning, and a high elemental sulfur recovery. ZnS concéntrate 1

^ \ r o.-^

First stage pressure leaching

i

S/L Separation

i o.-*-

1

i

t

Purifi catión

Second stage pressure leaching

''

''

Electrowinning

S/L Separation

\

Fíesidue ce ntainíng S

electrolyte Zinc

Figure 5.21 - Sherritt's two-stage pressure leaching technology for ZnS

When pressure leaching of zinc-lead concéntrate was conducted at 220°C, it was possible to solubilize zinc, copper, arsenic, and antimony sulfides and obtain a purified lead sulfate residue (Table 5.4) which was then treated by pyrometallurgical method. The leach solution was treated for copper and zinc recovery. The process was, however, shut down because SO^ was evolved from the blast furnace due to the decomposition of PbSO^.

115


Table 5.4 - Aqueous oxidation of Pb-Zn sulfide concéntrate by air and water at 220°C and 5500 kPa

Peed,%

Residue, %

Solution, g/L

50.6 8.7 6.5

51.2 1.2 0.8

Trace 49.5 47.8

Pb Zn Cu

Nickel and cobalt Nickel, cobalt, and copper are present together in different proportions either in matte or as precipitated sulfides. The matte is obtained by smelting of sulfide concentrates while the precipitated sulfides are obtained after leaching laterites by acid then precipitating the sulfides by HjS. In these cases a temperature of 200°C and an oxygen pressure of 4000 kPa are needed; the reaction is complete in 2-3 hours. Treatment of matte Nickel sulfide, NigSj, is obtained by smelting nickel sulfide or nickelcopper sulfide concentrates to form matte from which iron sulfide is then removed by oxidation and slagging. At Impala in South África, m Botswana in África, and in Germany the white metal is treated by oxygen in acid médium: Ni,S, + % 0 , + 2H" -^ 3Ni2^ + 2S0 2- + H,0 3

2

2

4

2

Cu^S + ^20^+ 2H"-> 2Cu2^+ S 0 / - + Hp Copper in solution is then precipitated by adding fresh white metal whereby more nickel sulfide is solubilized: NÍ3S2+ 3Cu2"-> 3Ni2^+ CuS + Cu^S

";^^

116


siiinniii TO «ECWEHt

««HE f«l« SHaTÍ!

Figure 5.23 - Leaching of nickel-cobalt matte

Chapter 5 - Leaching Processes in Presence of Oxygen

117

Leaching of precipitated sulfides. Nickel and cobalt sulfides obtained by precipitation from dilute solution by HjS were then shipped from Cuba to Port Nickel in Louisiana operated by Freeport Nickel to be solubilized at high temperature and pressure to get a solution suitable for electrolysis (Figure 5.24). When the plant in Cuba was nationalized, the sulfides were sent to the former Soviet Union for smelting. When the Soviet Union was disintegrated in 1990, the sulfides were shipped to Sherritt-Gordon in Alberta to be treated by hydrometallurgical method:

Ni-Cu matte

i_U

MS + O . ^ M S O ,

Leaching

NH,

ZE:

S/L Separation

where M = Ni and Co.

Iron removal O,

Pressure leaching

S/L Separation

F.,0.

Molybdenum When molybdenite, iVloSj, is leached in water in presence of oxygen, hydrolysis accompanies oxidation and as a result molybdic acid is formed as a white precipítate:

Evaporation S/L Separation

SO,

-i3

Purifi catión

CUjSe

Cuje '

MoS.-^Mo^"+2S2-

NiSO.

Solution

Mo^" + 2,Hp + 'ÁO^ -^ H^MoO^ + AW Platinum metáis concéntrate

S2-+20.

S/L Separation

I

Electrowinning

Copper

Figure 5.22 - Separation of copper and nickel from matte

Stillwater matte The Stillwater smelter in Montana produces matte containing 42% Ni, 27% Cu, 22.5%) S, and 2.1%) platinum group metáis. After leaching with dilute sulfuric acid and oxygen in autoclave the residue contains 60-65% Pt-concentrate.

S0,2-

Overall reaction: M0S2 + S H p + '/2O2 -^ HjMoO, + 2H2SO4

Rhenium associated with molybdenum in molybdenite is expected to be in solution. To prevent the formation of sulfuric acid, molybdenite can also be solubilized in an alkaline médium to form molybdate and sulfate:

> . T3

o

o

c3

03

nS -O

CíO O

-a co ^ 2 cd

'>.

1^

o¡

i-

ü

o3

o o

X

X co

+

CNJ

+ O o ^

CM

+ +

T

O

O

T n

+

+ 1

CN

^

^

T

CN

U)

o 03

+ CN

o

O o

a.

+

o ;^

_ 03

OH

03

=3

X

+ O

c o

C tu

o

o

2 o

o

•tí

i£ 13 -(-> (U

3

g^

c« _ , oi

r/1 1/3

CL

^ 8

q:; tu

-Td

os

^ -s a "^

o3

03

-t—>

03

00

03

c«

Ü

00

•^

CM

x;

t3

03

,P

^

CO

o

S

•S

Si

tu >

I/) ^

tu

o

u^ S o

tu DH

tu ^

O

u

U

C

1^

o a> P

C O

U) 3

> s tl> o x: o

Q

c •n CAÍ

i1>

03

e

tu X

!X

c/3 03

tu

tu

••"

00

B

3

P o

tu

>-§ tu

CLH

,0

(U

ii

IZl c/3

tu O

O

c/3 'JIÍ

o. 00 ' t í C 3

y ^ oo O)

§ e x; 2

S2

o u «

G •03

'a o *•

•s I

T3

^ o.

tu c/3

1^ P

oi

O

O

>tu 3 _(U ^ CG cC P "^

B 3 .5 «

c/3

-o c

t:; =! o 'B ^ o o Ô 2? 2 03 "^ ^ 1

T3 B

"T

CO u.

H ^

o

ex (N C/3

tu

>

tu

tu

o G

S "o

>.

tu

B S

î

o3 tu

03

>^

. c tu o

X

C

o S 3 tu -a 5 ts c 1/3 3 __. c O O

•-•

2 'ób

o t4-l

u -Td tí:;

03 (/) tu

X

O) (U

o

tu

OH

2 ^

tn 03
^

(L) :>

o

O (U •'£ -g o t-,

x; o

DO

« 131 ^

^ 2 .£ -S

•a

x; -t—»

-^

C 03 tu ,

X .S

+

3

3 :^

o -c

«

03

c«

c03

S «^ S

03

^^ s

-^ 3

o

CD

*-4-í

S ^

X) C

S

o

O

-O

O

(U XI

^s

03

o 5 2 o

X

O 03

I

^

(D Üfl " o

-a c

O

CO

O CD +

2

ofl

CN

X

O

"o g

CO CNJ

o CO

c

o, o

=3
O CVJ

CNJ

CM

0)

o

&^

oj

Q

CN

C

c •-

03

O

4>

5

pfi "•

1/3 ^.

X! tu c/3 3

" • c/3

00

K

CNI

+ CO CD

03

lU 03

§ G

O O — CJ 03 .in ^

c o

I>. o

'^

C

u

g

"cS tu c/3 03

ti
— X _3 O

3

c/i

[<19 f W ^ a ] P^^M w o j }E jjBqoo puB ]3?joiu JO ÁJ3A0D3J puB a m j u s o u o o spijins JO guTifOBaq - t/^-g s j n S í j 13SIBVH

Oi^

mmá wm 3mmi

OHMINIS 5=E.

si3nt)iaai3)ioiií

5tllii3nDI8B I—
13SHVN 01

ii30*od nvaoo

sa3iJissvi3

^

,H3iV3H3ad

SHIHSVÍ*

3A¥lD0inV

uniimS

Z^

mm

"¡ama 13Ü31N

ÍI013Í38 3dl
H S V U Y NOllOnOBS N3308!1AH 39(UiaiN33

aNiss30o«d snvs 3WJinS VIMOHNÍ

01 aonon

u 83Zinvi8180 •TOlVSOdW

> 1

mi Nfliimossio

«31113 ASVlOü

ísnss

Aüams samis 3gi3ifls

(S31HSN12)

Bm

SSHVl HDU¥lldia38d ISaU Oi 0310*33»

3iS¥» 01

3ñV130inV NOI13nfl38 N330MAH

mi

--yHs HSVT3

y-*

^solimos 31ÍV3 ' (S3IS3S NI s) mn aiHvi K0uyiic!i33ad

8311IJ 38(1SS3W

munumM

831113 { 3«nSS38d

33V801S 01

SnVS U7-5N \ 3101338 •• N0limoSlMVN33Bd

«oficn aN3 1881115 3013105

120 _ ressure Hydwmetallurpv

If, however, the reaction is conducted in presence of dilute acid (O.I M), the formation of elemental sulfur and ferric oxide takes place: 2FeS + ViOîaq)


Norilsk, Russia. 5FeS + 50, + 2H,0 -^ 4FeOOH + FeSO. + 4S

Fe203 + 2S

It will be also observed that the acid initially added will be present unchanged at the end of reaction. In fact the acid is consumed at the initial stage of the reaction and then regenerated later (Figure 5.25). It should be observed that the oxygen utilization in the presence of acid is less than in the first case. The reaction may be chemical or electrochemical in nature:

121

TL solutioH

'1

Pyrrhotite concéntrate pulp; TL solution Water from recycled supply systera

^^ ítem

Pulp after leaching Time

Time

Figure 5.25 - Aqueous oxidation of ferrous sulfíde or pyrrhotite. Leñ: in absence of acid; Right: in presence of acid

Figiire 5.26 - Leaching of nickel from pyrhotite at Norilsk, Russia Pvrrhotite concéntrate oxygen-enriched air ~

The oxidation of Fe^"" ion to Fe^'' and its hydrolysis follow the reactions:

Li

Aq. Oxidation I ron pellets Precipitation

2Fe2^+2H^+y20 ^2Fe''*+Hp Ttiickener

Fe^^ + SHÔ -> Fe(OH)3 + 3H" 2Fe(OH)3-^Fe203+3H20 It can be seen that in absence of acid the oxidation of ferrous ion to ferric is not possible. That is why FeS suspended in water will yield only a solution of FeSO^ when oxidized. In presence of acid, however, the formation of Fe^"" becomes possible, but because of hydrolysis, Fe203 is formed and the acid is regenerated. Thus, it seems as if the acid acted as a catalyst.

CaO + neplieline

Flotation

S°and suifides

Precipitation

Disintegrator 120°C

T

Flotation

pH8

T"

Tailing to disposal

Sulfide concéntrate to smelting

Figure 5.27 - Norilsk process

S"

Melting Sulfur 99.72%

- pTT I

122


Voisey Bay nickel sulfide process With the discovery of the Voisey Bay sulfide deposits in the Canadian North (Figure 5.28) it was decided in 2005 to use a pressure leaching process similar to that used for zinc and copper sulfides to produce elemental sulfur instead of SO^ (Figure 5.29). The decisión was made because the Government of Newfoundland refused that the concéntrate be shipped outside the Province. This is the first acid pressure leaching process for nickel sulfides. The sulfide concéntrate is mainly pendlandite-pyrrhotite. The mine is estimated to contain 141 million tonnes at 1.6% nickel. Voisey's Bay is a surface nickelcopper-cobalt mine located in northern Labrador, 37 km southwest of the town ofNain. The mine is operated by Vale Inco. Aplantis under construction to treat the sulfide concéntrate (Figure 5.30).

123


Pendlandite-Pvrrhotite concéntrate

u . r Aq. Oxidation

^2

1'

Filtration

i

Acid recycle

Purifi catión

1

L-

Recovery

—^

Flotation

Sulfur

1'

Concentration

Platinum metáis

t

Residue

Nickel

Figure 5.29 - Pressure leaching process for treating the Voisey Bay concéntrate Chemical Reagsnt Prepantion Iron Removal & Nautraltzation ..

Metal Separación & Purlficatlon (Solvant ExtracUon)

Solld/LIqukt Separatlon (CCO) >r*ssure Laaching Nielwl RacoMiy (eaelrcNvhmlng) Cruahing & Qrtnding Nickel Concéntrate Contelner»

Figure 5.30 - Nickel sulfide pressure leaching plant under construction at Argentia in Newfoundland

Argentia Froi»5Sirn Fag'lihf

Figure 5.28 - Location map of the nickel deposits in the Canadian North at Voisey Bay and the hydrometallurgical plant Argentia in Newfoundland Figure 5.31 - Pilot plant of nickel

TT' 124


('hapter 5 - Leaching Processes in Presence ofOxygen

Liberating of gold from pyrite and arsenopyrite Pyrite and arsenopyrite have received great attention recently because in some gold ores called "refractory", they entrap gold in their crystal structure and render the metal un-extractable by cyanide solution unless the mineral structure is destroyed by thermal or aqueous oxidation prior to cyanidation. At neutral pH, the formation of sulfate is favourable: FeSj + Hf) + ViO^ -^ Fe2^ + 2W + 2 S 0 / -

Aqueous oxidation at high temperature and pressure in alkaline médium yield FGJOJ and sulfate ions: 2FeS2+y202+80H-

Fe203 + 4 S O / - + 4 H p

The advantage of such reaction is no corrosión problems but the disadvantage is the high cost of the reagent. The recovery of gold from pyrite by leaching at high temperature and pressure in neutral or acid conditions is used world wide since 1985 (Table 5.5). Brazil AngloGold Ashanti Brasil (AGA Brasil) has commissioned in 2012 a refractory gold pressure oxidation plant at the site of the original Sao Bento operation in Mina Gerais, Brazil which was operated between 1986 and 2007. When the ore body was exhausted AGA Brasil purchased the facility from Eldorado Gold, to process refractory gold concentrates produced from the nearby Córrego do Sitio mining área. The original autoclaves were limited to a máximum operating temperature of 190°C and a pressure of 1700 kPa but were of sufficient size to allow extended pressure oxidation retention times for treatment of the new concéntrate (Figures 5.32 and 5.33). Antimony and arsenic were precipitated to near completion in the first stage of neutralization to a pH of 3 to 5.

125

Table 5.5 - Liberation of gold from pyrite and arsenopyrite Start up 1985

Plant Location

McLaughIin USA 1986* San Bento Brazil 1988 Mercur, Utah USA 1989 Getchell USA 1990 Goldstrike Nevada, USA 1991 Goldstrike Nevada, USA 1991 Porgera, Papua New Guinea 1991 Campbell Canadá 1992 Lihir, Papua New Guinea 1993 Goldstrike USA 1994 Porgera, Papua New Guinea 1997 Lihir, Papua New Guinea 1999 Twin Creeks, Nevada, USA 1999 Macraes, New Zealand 2006 Madang, Papua New Guinea 2009 Kittila, Finland 2009 Pueblo Viejo, Dominican Republic 2012 Petropaviovsk, Amur región, Russia 2012 Sao Bento Brazil

Owner Homestake USA Genmin S. África * American Barrick Canadá First Miss Gold American Barrick Canadá American Barrick Canadá Placer Dome Canadá** Placer Dome Canadá ** Nerco Minerals

Feed

Capacíty t/d

Number of autoclaves 3

ore

2700

concéntrate

240

2

ore

680

1

ore

2730

3

ore

1360

1

ore

5450

3

concéntrate

1350

3

concéntrate

70

1

concéntrate

90

1

ore

11580

6

American Barrick Canadá Placer Dome Canadá ** Rio Tinto

concéntrate

2700

6

—

—

3

Newmont

concéntrate

Macraes Goldfield

concéntrate

1

Highiands Pacific Agnico-Eagle Barrick

concéntrate ore

Polymetal International

concéntrate

AngloGold Ashanti Brasil

6000

1 4 The worids largest 6

T 126


t hapter 5 - Leaching Processes in Presence of Oxygen

North Atlantic

127

Ocean

MonteCristi^ P u e r t o Plata Dajabon M9^ Sabaneta

.Moca

San Juan

DOMINICAN REPUBLIC

:iagp"'

Salcedo La Vega,_,

lasPiita

i SanF' ÜeMa

Co.tu ©• - ^ ' BorV.v

-ilLLEF. ,ITRAL

- "^gua

¡.>e Samaná á

co

de 1.a Mar ^•r Plata ^Viato Mayor

\ j

Figure 5.32 - Location map of AngloGold Ashanti Brazil

SANTO DOMINGO

SanP. deâ^c

El Seibo Cape - H i g ü e y Eingaño

I Romana

San C r . i l ó b a í

Las Lagunas

Oxygen

Concéntrate

i r-

Sulphuric Acid ' \ ^

\ S

Acidulation L

Quench Water

^

\

Lime; Cyanide

^—

Pressure Oxidation

CCD Wash k

f ümestone Lime

'' Solution Neutraüzation

1, \ ^

L

S

Cyanide Leacli

B e a t a 1,.

• Gold

jpe Beata

Caribbean

Sea

A l t o Velo 1,

1 ^ Tailings to Impoundment

Figure 5.34 - Location map of Pueblo Viejo and Las Lagunes

\ v

1

1

Figure 5.33 - Pressure leaching plant at AngloGold Ashanti Brazil

Dominican Republic The $3.8 billion Pueblo Viejo mine (Figure 5.34) owned by Barrick Gold and Goldcorp holds 25 million ounces of proven and probable reserves. There are four autoclaves each is 6 m diameter and 40 m long and processes approximately 6000 tonne ore/day liberating about one million ounces of gold per year (Figures 5.35 and 5.36). The autoclaves opérate at 230°C and 3450 kPa and residence 60 to 75 minutes. The autoclaves are the world's largest brick-lined autoclaves.

Figure 5.35 - View inside the autoclaves hall

Mona

129


iion tonnes of ore with a grade of 4.8g of gold per ton. The majority of the gold is found in arsenopyrite and pyrite and a small fraction in the outer oxidized portions of the minerals. The ore from Kittila is processed through flotation, pressure oxidation, and carbon-in-leach circuits, and electrowinning (Figure 5.38).

|-5^

/

Sweden (

S

Norway N

\

Oslo

S

Finland

/ Stockholrn

Helsinki

Tailings from the Pueblo Viejo mine derived from operations between 1992 and 1999 and stock piled at Las Lagunas are about 5 million tons grading 3.8g/t gold and 38.6 g/t silver. They are being re-processed through ñotation, foUowed by sulfide oxidation using the Albion process prior to extraction of gold and silver utilizing standard carbon-in-leach cyanidation. The Albion process involves fine grinding then leaching at atmospheric pressure. PanTerra Gold anticipates annual production of 69,000 oz Au and 630,000 oz Ag. The first gold was produced in 2012. Finland The Kittila gold mine is located 900 km to the north of Helsinki at the Suurikuusikko gold deposit (Figure 5.37). The mine is one of the largest gold-producing mines in Europe. Its commercial production began in May 2009. It is operated by Agnico-Eagle Mines and is expected to produce an average of 173,000 oz of gold a year and has an estimated lifespan of 15 years. The mine contains an estimated 4 million oz of proven gold reserves. The reserves consist of 26 mil-

Denmark ^-.^5-'^ /

i0

\

SL Pecersburg

•

Russia ^^ \

Figure 5.36 - View of one of the autoclaves

y-

• )\ Estonia

r^ v^Tx

Figure 5.37 - Location map of Kittila gold mine in Finland

r~i

n

Crushing

1/

pit*

Cíirbon Flotation

Sufphur Ftotation

' — - — .

Umtargrouml

I

^^^H

C.I.L Tatling Di&posal

Leaching ÍC!. ' x~r\

nf Flotation TatJíng Dl&posal

-g Cyanide Dostfijcno

piH-'

-{jr^h-

sa

—'

Docó Bar

l'igure 5.38 - Flowsheet of Kittila aqueous oxidation of gold ore under pressure

130



131

Russia

By the end of 2013, Polymetal International will start treating its refractory gold ores at the Pokrovka mine [Malomir and Pioneer deposits] in the Russian Far East (Figure 5.39) by pressure oxidation process (Figure 5.40). The concéntrate, ground to 90% -44|a.m, in the form of a pulp is fed to the acid treatment facilities, where carbonates are decomposed. The acidic pulp is pumped into a threestage counter-current washer to decrease the chloride concentration to a mínimum. The thickened product is then fed into the autoclave using a high-pressure pump. Pressure oxidation is carried out in a horizontal autoclave at 225-23 0°C and an oxygen pressure of 0.5-0.7 Mpa, with the total pressure in the autoclave at 3.2-3.5M Pa.

Figure 5.40 - Flowsheet of leaching of refractory gold ore. 1. Tank for acidic treatment, 2. Thickeners for chloride washing, 3. Autoclave feed tank, 4. High pressure pump, 5. Autoclave, 6. Flash tank, 7. Scrubber, 8. Conditioning tanks, 9. Thickener, 10. Filter press, 11. Solution neutralization tank, 12. Thickener

The autoclave is divided into four sections, the first of which is twice the síze of the others. The autoclave also has five impellers, of which two are located in the first section. Oxygen is supplied from the underside of each impeller and cooling water is supplied to each section independently. The oxidation of pyrite and arsenopyrite are exothermic, and enables the process to run auto-thermaliy. Any excessive heat is controlled by feeding cold water into the autoclave.

Figure 5.39 - Location map of the gold mines in the Amur región, Russia. 1.Pioneer mine, 2. Pokrovskiy, 3. Malomir, 4. Albyn

Oxidized pulp from the last section of the autoclave is discharged into two flash tanks connected in a series. The pressure in the first tank is 0.7 Mpa, with the pulp temperature at 170°C, with the second unit at atmospheric pressure with a pulp temperature of ~100°C. The pulp is thickened and filtered then the cake is washed and delivered for cyanidation. The acidic autoclave solution is then neutralized,

132



High-grade oxide and sulfide ore is treated by milling and cyanidation, but for the lower-grade oxides heap-leaching is used. For the refractory ores the company completed in 1994 a roasting unit for higher grades and a bio-oxidation system for lower grades. Refractory ore with a carbonaceous content is treated in the bio facility or by ammonium thiosulfate leaching. The Winnemucca operations use autoclaves to pre-treat refractory ores. In 2005, open pits mined 175 million tonnes of material and the underground mines 1.42 million tonnes. The oxide milis processed 4.20 Mt averaging 4.3g/t gold, the refractory milis 8.15 million tonnes averaging 6.8g/t, and leach dumps 17.5 million tonnes averaging 0.9g/t to give a total output of 2.46 million oz of gold.

first with limestone to a pH of 4.5, and then with lime, to a pH of 9-10, before it is pumped to the tailings facility. The key factor affecting the gold recovery from the carbon-bearing Malomir concéntrate was the presence of chloride ions in the process water. In 2011, a pilot pressure oxidation unit was installed at the Petropavlovsk pilot plant in Blagoveschensk. To process the flotation concentrates from both the Pioneer and Malomir deposits, it was decided to build a centralized pressure-oxidation unit at the Pokrovskiy mine, 670 km from Malomir and 40km from Pioneer. The location was chosen for its existing supporting infrastructure, including a resin-in-pulp process plant.

Barrick Gold already has a pressure leaching plant in Elko, Nevada Ibr liberating gold in pyrite followed by cyanidation (Figure 5.42).

Outotec in Finland will manage the design and construction of the pressure oxidation plant in collaboration with Gidrometallurgiya R&D Centre in Saint Petersburg. It will supply 6 horizontal autoclaves each of 3m diameter and 15m long, a total capacity of 90 m3 and an operational capacity of 50 m3. The acid-proof brick lining will be produced and inserted by DSB in Germany. USA Newmont Mining Corporation started mining gold at Carlin, Nevada in 1965 (Figure 5.41). In 2001, it acquired Battle Mountain Gold and in 2002 Normandy Mining in Nevada. Most operations are located on the Carlin Trend west of Elko. The Twin Creeks and Lone Tree Complex are in the Winnemucca región further west. The Phoenix gold-copper project near Battle Mountain produces about 420,000oz/y of gold and 21,000 t/yofcopper.

133

InOopondance group

_Elko Ctrlln tnna

€^ Figure 5.42 - Barrick Gold pressure leaching plant for liberating gold in pyrite at Elko

Figure 5.41 - Carlin Trend in Nevada

k.

New Zealand The Macraes Goldfield is New Zealand's largest gold producing operation, consisting of the Macraes Open Pit and Frasers Underground mine. The Macraes mine has been in operation since 1990 and produces about 130 000 oz/annum. Macraes is located 100 km

134



north of Dunedin in the Otago región of the South Island of New Zealand (Figure 5.43) The Frasers Underground was commissioned in 2008. The processing plant is situated within short distance of the Macraes Open Pit and includes a pressure oxidation plant for the processing of sulfide ore, carbón in leach, and elctrowinning (Figure 5.44). Refractory concéntrate from the Reefton processing plant is transported by road and rail to Macraes.

NorthlaníK

135

JUjrOOJM aX tHCXEKRS

*MyM^*gvt^,___

^^

^^

AuckiandV^^ l ^ Y"

TaranakiC

,,.|Bay of Plenty

/--'••:, ^ < ^

/

y

ftUWMÍlUMwa

/ Hawkés Bay

Reefton Open Pit

fUaROMHNMG

Figure 5,44 - Macraes gold processing plant

SELENIDES AND TELLURIDES FROM ANODIC SLIMES y ,.

y^Christchurch

Introduction Macraes Open Pit Frasers Underground

^ 'K,., ^ ^

jY p>

New Zealand Dunedin

Southlánd CK

Figure 5.43 - New Zealand's gold mines

The Reefton mine was commissioned in 2007. A gold bearing concéntrate is produced at the mine which is then railed over 600 km south to Palmerston from where it is trucked to the Macraes operation for processing. Reefton produced 85,843 ounces of gold in 2010.

Selenium and tellurium are mostly in association with non-ferrous metal sulfides, especially those of copper and nickel. During pyrometallurgical processing of concentrates of these metáis, appreciable amounts of selenium and tellurium are volatilized. The remainder deposits during the electrolytic refining as slimes at the bottom of the cell. For example, fire-refined copper contains 0.01-0.02% Se and up to 0.004% Te, as selenides and tellurides of gold, silver, and copper. The slimes are a grayish black powder, minus 200 mesh. From 2 to 20 kg of slimes are produced per ton of copper cathode; they are usually collected every 14 to 21 days of electrolysis. A typical analysis of slimes at a refinery in Canadá is given in Table 5.6.

T 136


Table 5.6 - Typical analysis of anodic slimes at the Canadian Copper Refiners, Montreal East %

% Cu Ag Au Se Te Pb

30 21 1 15 5.5 10

As Sb Bi Sn Si Balance*

0.25 1 0.3 0.5 1 14.45


copper sulfate solution is evaporated to crystallize CuSO^'SHjO for the market. The residue from pressure leaching containing mainly selenium, silver selenide, gold, and lead sulfate is pelletized, then roasted at 815°C in air. Selenium dioxide is recovered in the scrubbers, and the precious metal fraction collected from the roaster is melted in the usual way in a doré furnace. Slimes

•TLj_r Filtration ir Solids Drying

Pressure leaching takes place with dilute HjSO^ (30%) and oxygen under pressure are used. The process is conducted in stainless steel autoclaves at 125°C and 300 kPa and the reactions taking place are the following:

H 0

1

Scrubbers M— SO,

"i

,' "

C u j e + 2H^+ 5/2O2 ^2Cu2" + T e O / ' + H p

Cu 1

1

Acid process

2Cu2^+Se + 2 K O

H.,SO.

Pressure leaching

•IVlainly sulfur and oxygen (as PbSO^, CuSO,), SiO^, and traces of Fe, Ni, Al, Ca, and Mg.

Cu2Se + 4 H " + 0 2

137

Filtration

1

Ir

—•Cementation

Ai

1

T

Leaching

1 t

Filtration

f

1 CuSO^- 5H2O

°^ - ^ 1 Cu Je

Air - •

Oxidation Melting

Crystallization

^

t Filtration H,S

Pelletization

Solution

—' CuSO^ solution

1

Slag

oíd and silver

Selenium

Cu + 2W+ 'ÁO^^ Cu2"+ Hp

Under the leaching conditions selenium is precipitated in the elemental form, while the tellurium goes into solution. A flowsheet of this process is shown in Figure 5.45. After the solid-liquid separation step, the solution containing copper and tellurium is agitated with metallic copper in form of pellets to precipítate tellurium as CUjTe: TeO/-+ 5Cu + 8 H " ^ C u j e + 3Cu2"+ 4H2O

Excess copper is added to neutralize the remaining acid, then the

Figures .45 - Pressure leaching of anodic slimes at Canadian Copper Refiners, Montreal East, Canadá

In a recent development, the steps comprising pressure leaching and air oxidation of residue are replaced by a single high-temperature oxidation in a top-blown rotary converter to get directly doré metal. Gases evolved during this treatment are collected for selenium and tellurium recovery. Multi-stage leaching process This process (Figure 5.46) was developed in Finland by Outokumpu Company. The slimes are not filtered but treated directly as a slurry in the electrolyte. The process involves the following steps:

138


H.SO,

Anodic slimes Air u

ir

u

[^^1

L^H

Leaching 80°C

i

Filtration H "íO " 2 ^ "-'4

^|

1

Cementation

±

Solution

Alkaline process Sodium hydroxide and oxygen under 1400 kPa pressure decompose

1^ Leaching 160°C " Filtration

I

'

NiSO,

Cuje

SO,-^

which is collected by filtration and refined by distillation. • Formation of doré metal. The residue from the previous step is melted with fluxes to form doré metal which is refined electrolytically.

-•-CuSO^ solution

•

Cu

139


selenides and tellurides of copper and silver at 150°C as follows:

"2^ rjasfíc

T

Heatinc Melting 1

i

H,SeO, +Se

Doré metal

Figure 5.46 - Outokumpu Process for the treatment of anodic slimes

• Leaching of copper. Metalhc copper in the sHmes is first removed selectively by air oxidation at 80°C: Cu + 'AO^ + 2H" -^ Cu2^ + Hp

^^^1 ^^^H

CUjSe + 2O2+ 2 0 H - ^ 2CuO + SeOg^^Hp CUjTe + 2O2+ 2 0 H - ^ 2CuO + TeOg^-n Hp Ag^Se + 3/2O2 + 20H--^ A g p + SeOg^^Hp A g j e + 3/2O2 + 20H- -^ A g p + TeOg^- + HÔ The selenites and tellurites are soluble in the solution. However, further oxidation converts all the tellurite to insoluble sodium tellurate: Na2Te03+ VzO^^ NaJeO^

• Leaching of nickel and tellurium. Nickel oxide in the slimes is removed by dilute H2SO4 at 160°C in pressure reactors:

Anodic slimes

NiO + 2H"^NP^+H20 NaOH-

In this operation most of the tellurium and any remaining copper go into solution. After filtration, tellurium is precipitated from solution by cementation with metallic copper. • Selenium recovery. Selenides were found to decompose readily at 600°C in an atmosphere of SOjí Ag,Se + S O2(g), , ^ S e O2(g) , +AgS Volatilized Se02 is captured in the gas scrubbing system. Due to the presence of SOj in solution, elemental selenium is formed

-02

Y

T

Y HESO.,

Leaching

Filtration

Seienium recovery

, Solids \^~>\

Filtration

.Solids |-

Au, Ag ""recovery

Te recovery

Se

Figure 5.47 - Sodium hydroxide process for treatment of anodic slimes from electroiytic copper refining.

140


Only a small amount of selenite is oxidized to insoluble selenate. The slurry is filtered and selenium is recovered from the sodium selenite solution, while the residue is treated further with H2SO4 to dissolve the tellurates. Gold and silver are recovered from the insoluble residue (Figure 5.47).


141

humidity from the atmosphere. Proper storage of these materials is therefore of special concern. Some arsenic compounds are used as insecticides, weed killers, and wood preservative.

ARSENIDES Arsenic compounds are highly poisonous and therefore treatment of arsenic ores needs special measures to protect the workers and the environment. The most important arsenic minerals are shown in Table 5.7. Arsenical ores may be treated in two ways: • Leaching with acid or alkali in presence of an oxidizing agent. • Melting in presence of fluxes to volatilize as much as arsenic and sulfur as possible, and the resulting product, called speiss, is then leached. A speiss is mainly a complex mixture of metal arsenides whose composition varies widely, depending on the type of ore treated. A typical speiss contains 15-35% arsenic.

10

12

14

Figure 5.48 -Distribution oftrivalent arsenic species as a fünction of pH 100

Table 5.7 - Arsenic minerals Arsenides

Sulfarsenides

Arsenic sulfides

Niccolite Smaltite Sl
NiAs C0AS2 C0AS3

Cobaltite Enargite Arsenopyrite

CoAsS

Realgar Orpiment

AS4S4

-2

o

2

4

6 pH

8

10

12

14

Figure 5.49 - Distribution of penta valent arsenic species as a fünction of pH

CU3ASS4

FeAsS

AS2S3

Arsenic forms two oxides AsÔg and ASjO^ which dissolve in water to form a variety of species, depending on the pH, as shown in Figures 5.48 and 5.49 respectively. Arsenious oxide is volatile while arsenic oxide is not volatile. Arsine, AsHj, is a highly poisonous gas that forms from arsenides by the action of dilute acids or even

Arsenide ores and concentrates can be leached by acids or alkalies. In the first case both the metal and arsenic go into solution while in the second case only arsenic is solubilized. Acid leaching This process (Figure 5.50) has the advantage of eliminating the danger of having arsenic in solution; instead, it is removed as an insoluble residue. The process has been applied for the recovery of


142


143

cobalt from a sulfarsenide ore. Leaching is carried out at 200°C, and under oxygen pressure of about 1200 kPa. Metals are converted to soluble sulfates while iron and arsenic salts are oxidized, and these combine to form insoluble iron arsenate: Asenic ore Air

H2SO4

Y

V

V

Pressure leaching

Fiitration

-CaS04, FeAs04

NH3-

Y

Y Ammlne formation

H2-

i

Cobalt precipifation

•a

o a U

Fillration

I

Cobalt

Figure 5:50 - Calera Process for pressure leaching of arsenical ores

CoAsS + ViO^ + Hp -^ Co^^ + SO42-+ As04^+ 2H"

The reaction is slow unless dilute acid is present. The slurry is filtered to recovar cobalt from the solution. The process is known as the Calera process (figure 5.51). Analysis of the concéntrate processed is given in Table 5.8.

3 M


144


Table 5.8 - Analysis of cobalt-arsenic sulfide concéntrate at Calera Cobalt Refinery

The residue can then be processed by conventional methods to yield cobalt (or nickel) free from arsenic. Treatment is carried out at 115°C, 840 kPa oxygen pressure. Arsenic can be removed from the Icach solution by precipitation with lime to give calcium arsenate, or with H2S to give arsenic sulfide. When lime is used, sodium hydroxide is regenerated:

% 24 17.5 0.5 20 1 29 5 12

As Co Cu Fe Ni S Gangue H2O

í

2 A S 0 / - + 3 C a ( 0 H ) , ^ Ca3(AsO,)2 + 6 0 H -

The process is known as the Sill Process after its inventor Harley .Sill. Analysis of an ore treated by this process is given in Table 5.9. Table 5.9 -Analysis of ore processed by sodium hydroxide leaching.

Alkaline leaching Oxidation in presence of NaOH results in solubilizing both arsenic and sulfur to arsenate and sulfate respectively, while the metal valúes remain un-dissolved (Figure 5.52):

%

Ore NaOH

Y t Y

Leaching 800kPa,115C

i

CaO Solu ion

Filtration •^ y Residíje{0 1% Leaching

t

(gangue, Ag)

, NajSOí Arsenic — > | Crystaircation | - > precipitation

Calcium arsenate

Filtration ySolutií )n

i

As

45

S

20

Fe

19

Co

12

Ni

3

Cu

1

Ag

100-150 oz/ton

III Russia, ammonia is used for ore containing 1.2% Co, 1.5% Ni, .ind 7% As (Figure 5.53). The ammoniacal solution containing cobalt and nickel is distilled to recover ammonia and precipítate a concéntrate containing 22% Co and 30% Ni. It is further treated to gct a concéntrate containing 70% Ni + Co.

Purification

i

i

Precipitation Miu 10 m m%m -•

I, ( c o , 10 AÍSMIED

Y Co(OH)3

Figure 5.52 - Leaching of arsenide ore with sodium hydroxide DISTIUATiDH

V^

COLUHN

("I—ZD—1

C0ASS + 7/2O2+5OH-

Co(OH)3+ S 0 / - + As04^+ Hp

145

•

Co-m-Ai

mzmmi

1

In T

Ev*?aiuioii

ntiut

Figure 5.53 - High pressure leaching of cobalt-arsenic ore



146

but half of the diagonal positions are occupied by pairs of the sulfur atoms and the other half by pairs of arsenic atoms. Marcasite is another crystalline form of pyrite in which the iron atoms occupy a body-centered cubic lattice with pairs of sulfur atoms on the diagonals.

Arsenopyrite Arsenopyrite may dissociate in water as follows: 2FeAsS (s)

147

2Fe^^3„-Âs-+S-

The behavior of disulfide ion is probably the same as in the case of pyrite in acid and in neutral médium, while the diarsenide ion forms arsenic acid:

UNSUCCESSFUL PRESSURE LEACHING PROCESSES

AS/-+ 30^ + 2H* + 2H2O -^ 2H3ASO,

Unsuccessful pressure leaching processes operated in the 1970s include the Clear process, Sherritt-Cominco process, and LurgiMitterberg process.

The overall reactions are: • In acid médium: 4FeAsS + 7O2 + 8H" + 2Hp -^ 4H3ASO4 + A¥é^* + 4S

Clear process

• In neutral médium:

The Duval Corporation in Tucson, Arizona was producing 40,000 tons/year copper from a chalcopyrite concéntrate at its Suharita plant in 1976-1983 by a pressure hydrometallurgical process known as the CEAR Process, an acronym for the process steps: CopperLeach-Electrolysis-And-Regeneration (Figure 5.54).

4FeAsS + 13O2 + 6H2O ^ 4H3ASO4 + 4Fe2" + 4SO,2-

• In alkaline médium: 2FeAsS + 10OH-+ 7 0 ^ ^ FeÔj + 2As04^ + 2 S 0 / - + bHp

^ ^, ^ „, CuClj, FeClj

Ferrous ion formed in the above reactions oxidizes further to form ferric arsenate precipitates:

Chalcopyrite '

1

V

NaCI + KCI

.

Leaching stage 1

"

2Fe2^ + 'ÁO^ + 2W -^ 2Fe3" + Hp

Filtration

°-'^\

Fe3^+ H,AsO + 2H,0 -^ FeAsO + 2H-0 + 3H" 3

4

2

4

¿

Pyrite and arsenopyrite are different from other sulfide minerals since they contain the disulfide ion, S^'\ arsenopyrite contains in addition the diarsenide ion, As2^". In pyrite, FeSj, the iron atoms are in a face-centered cubic arrangement with pairs of the sulfur atoms located on the cube diagonals. In arsenopyrite, FeAsS, the iron atoms are also in a face-centered cubic arrangement like in pyrite

011^1

'

Ff^CI

Leaching stage 2

olysis

^^

V

stion

t

Residue S, FeOOH, jarosite, gangue

ü

Washing

H,0

i

Melting, casting

" Electrore fining Cu

Figure 5.54 - CLEAR process

Ag


148

The product is described a "blister grade copper crystals". The process made use of the valency change of both the copper and iron ions in solution. At no point in the circuit, an acid was added; however, the Solutions were saturated with sodium and potassium chlorides not only to keep CuCI in solution, but also to increase the rate of leaching. A mixture of CUCI2 and FeClg was used as leaching agent, but the leaching action was predominantly due to Cu^"^ ion. Chalcopyrite is only slightly soluble in CUCI2 solution even at 120°C but, in presence of excess Cl" ion dissolution readily takes place because of shifting the reaction to right due to complex formation for both the anodic and cathodic reactions as follows:


149

is 30 m long and about 5 m diameter, divided into 10 compartments each fitted with a powerful turbine agitator. The autoclave is not insulated; the heat lost by radiation is compensated by that generated during the reaction. Copper in solution leaving the autoclave is mainly in the cuprous state. Use is made of this fact in the subsequent electrowinning step. That is why the solutions are handled carefuUy without excessive exposure to the air to avoid oxidation. Also, after cooling to 80°C and settling in closed thickeners, some cement copper is added with continuous agitation to reduce the remaining cupric ion: Cu2*+Cu-^2Cu*

Anodic: C u F e S ^ ^ Cu* + Fe^* + 2S + 3e-

Cathodic:Cu2*+e-^Cu* Cu* + nCI- -> CuCI^("-^>The process can be summarized as follows: First stage leaching In this stage the molar ratio Cu^*/ Fe^* in the leach solution is about 5, and only oné half of the chalcopyrite is solubilized according to the following reactions: CuFeS^ + 3CUCI2

4CuCI + FeCl2+2S

CuFeSj+SFeCIj-

CuCI + 4FeCI + 2S

The molar ratio CIVCuCI necessary to keep CuCI in solution and to yield reasonable reaction rate is about 20. The chloride ion is added as NaCI and KCI. Leaching is conducted at 105-110°C in a horizontal autoclave made of fibreglass (Figure 5.55). The autoclave

Figure 5.55 - Fiber glass autoclaves

The solution is then sent to the electrolytic cells for copper recovery, while the residue is sent to the second stage leaching. Second stage leaching In this stage the remaining un-reacted chalcopyrite is leached in presence of oxygen at 600 kPa, and 135-140°C for 90 minutes in a titanium autoclave. The leach solution is spent electrolyte in which the molar ration Cu^VFe^* is 10 (as compared to 5 in the first stage). In addition, a great part of the iron however, is in the ferrous state.


150

Although the excess chloride ion is not needed as a complexing agent, nevertheless, the solution is saturated with NaCI and KCI, since these are recycle solutions. In this stage, not only chalcopyrite is solubilized as in the first stage, but also both the ferrous and the cuprous ions are oxidized further due to the presence oí oxygen:


151

bottom of the cell. The reactions taking place are the following (ignoring the complexing Cl' ion): Cathodic:Cu*+e--^Cu Anodic:Cu"^Cu2^+e-

Ife^'+lW+VzO,

2Fe3^+H20

Overall reaction: 2 C u " ^ Cu + Cu^* 2Cu^+2H"+y20,

2Cu2"+H20

Although no acid is added in the system, the hydrolysis and precipitation of ferric ion furnishes the acid necessary for oxidation: 2Fe3^+ 4H2O -^ 2FeOOH + 6H^

Thus, the overall reaction that takes place is: 2Fe2^ + 4Cu" + 1 YzO^ + Up -^ 2FeOOH + AOu"' It is not essential to precipítate all the ferric ion since the solutions will be recycled to the first stage later. However, due to the oxidation of some of the elemental sulfur, a minor amount of sulfate ions are present in solution. This leads to the precipitation of some basic iron sulfate in the form of potassium jarosite: 3Fe3^+ 280^2-+ ^.+ e O H " ^ KFejíSOJ^ÍOH)^

which is suitable as a soil conditioner. It contains 20-25% S in the elemental form and < 0.5% Cu as chalcopyrite. Electrolysis Copper is recovered from the first stage leach solution by electrolysis in a specially designed diaphragm cells, using a copper cathode and a graphite anode. Copper is deposited in form of large dendrides which is scraped continuously by mechanical conveyers from the

It can be seen that only half of the copper is precipitated while the other half is recycled as a leaching agent in form of CUCI2. About 10% of the ferrous ion is oxidized to ferric at the anode. A small amount of chlorine is also evolved. Copper produced by this process is 99.9+%, but still not suitable for the market because it contains about 24 ppm silver (also about 100 ppm Fe). No method was found to remove silver from the solution before the electrolytic step. Refining The product is described a "blister grade copper crystals". As a result of the presence of silver in the product the copper powder is melted and cast in form of anodes for electrorefining to recover the silver and have a product acceptable for the market. Remarks For each tonne copper produced, 9 tonnes of salt (KCI + NaCI) are in continuous circulation. The effort done to keep copper in solution in the cuprous form is not after all exploited in the electrowinning step, because only half of the copper is precipitated in the elemental form and the other half is oxidized to the cupric state, i.e., the recovery step is essentially a disproportionation reaction: 2Cu^ ^ Cu + Cu2^ which may be conducted more efficiently and possibly at a lower cost using other methods if at all necessary since in this case cementa-


154

Chapter 6 - Precipitation

c o + Hp -> CO2 + :2H"+ 2e-

1

SO2 + Hp -^ H2SO3 -^ 2H^ + SO32-

•

s o 2- + H,0 -^ s o 2- -^ 3

2

2H' +

4

Oxides may also be precipitated in this way, While precipitation by hydrogen and carbón monoxide are non-ionic, precipitation by sulfur dioxide is ionic. PRECIPITATION \) Reduction

i

ByH2 Nickel Cobalt VOi

1

ByS02

By H2S

\

1

Silver Copper

Copper

Nis CoS

H H

\

Precipitation by hydrogen

Nickel and cobalt are precipitated on industrial scale by this method. ^ ^ H Copper can also be precipitated but not industrially applied. Reduc^ ^ H tion is usually carried out in horizontal stainless steel autoclaves equipped with agitators, baffles, heating or cooling coils, and the necessary connections for feed and gas inlets and outlets. The prod1 uct of this technique is a high-purity powder that can be used as such, or in case of metáis, hot pressed and rolled in form of strips. Precipitation may be conducted from aqueous as well as from nonH aqueous media.

ByCO

J\

PRECIPITATION BY REDUCTION

^ H Nickel and cobalt

\

Ionio

155

•

Theoretical basis For the reaction:

Figure 6.1 - Precipitation under pressure

the equilibrium constant is given by: Precipitation by hydrogen sulfide, on the other hand is ionic. It is based on the fact that when a H2S is added to a solution containing metal ion, a sulfide is formed whose solubility is very low under these conditions that precipitation takes place immediately:

K =

[Wf [M2^] •

PH,

Therefore: M2- + 32- -^ MS log[M2^] = -2pH-(logK + logP^2)

While CuS precipitates at ambient conditions, NiS and CoS presipitate at high temperature and pressure and in present of a catalyst. That is why autoclaves are used in this case. Precipitation of iron oxide by hydrolysis is also an important topic in pressure hydrometallurgy and will be discussed later.

This means that when precipitation is carried out at constant hydrogen pressure and constant temperature, then at equilibrium there is a linear relation between logfM^*] and the pH of the solution, and the slope of this straight line equals -2. This was confirmed for the precipitation of nickel from NiSO^ solution, as shown in Figure 6.2 and for cobalt as shown in Figure 6.3.

156


10.0

157


copper, nickel, and cobalt, this is conveniently done by operating in ammoniacal médium: [M(NH3)J2- ^ nNHg + M2+ M2^ + H^ ^ M + 2H^ H" + NH, -^ N H / 3

4

It can be seen that increasing the ammonia concentration has two opposing effects: 1

2 Equilibrium pH

Figure 6.2 - Precipitation of nickel from nickel sulfate solution by hydrogen at 3,500 kPa, [(NH4)2SOJ = 112 g/L, equilibrium conditions

• Precipitation is favored due to the neutralization of the liberated acid. • Precipitation is hindered because of the decrease in the reducible metal ions M^* due to the complexing action. Therefore, there must be an optimum [NH3] / [M^"^] ratio at which these opposing effects are balanced. In the precipitation of nickel, the optimum molar ratio was found to equal two, which agrees with the overall reaction: W* + 2NH3 + H2 l

i

^ 40 L _i

30

2.0 3.0 Equilibrium pH

4.0

Figure 6.3 - Precipitation of cobalt and nickel from acid solution, temp. 190°C, H pressure 3500 kPa, [(NHJ^SOJ = 112 g/L

<.-Û lo ce 10

l

í

/ ^ 0 ^^">5^

/

01

1.0

i

0

x:

7

l

NÍ + 2NH.

' P/

~

^'"^^

P

^*\^

^^\.. ^""^^^

/

^"^^.^

l

0

1

i

2

l

3

i

l

4

í

5

6

Molar ratio [NH3] / [Co^+j

It is olear from the above equation that more metal will be deposited if the hydrogen ions are removed as soon as they are formed. For

Figure 6.4 - Effect of the molar ratio [NH3] / [Co^*] on the rate of precipitation; temp. 200°C, catalyst H^PtCig 5.8x10-5 ^^ |_|^ pressure 3000 kPa

Pressure Hydrom etallurgy

755

The amount of ammonia in solution also influences the rate of precipitation. In the case of cobalt, the rate of precipitation achieves a máximum when [NH3] / [Co^*] = 2, as shown in Figure 6.4. Similar results were also reported for copper. Another way of removing the hydrogen ions as soon as they are formed during reduction is by reducing hydroxide slurries: M(OH), ^ W^ + 20HM2^ + H2 - ^ M + 2H"

OH- + H" ^ H,0 Overall reaction: M(0H)2 + H2

M + Hp

For nickel and cobalt, it was found that this reaction takes place at 270°C which is a much higher temperature than normally used, but the product is of extremely fine particle size. Nucleation In some cases, the presence of a solid surface for precipitation is essential; such a solid is termed a catalyst. Strictly speaking, the process is heterogeneous (contact catalysis) but is different from the heterogeneous process described later. If no catalyst were provided, the internal surface of the autoclave itself acts as a catalyst and deposition of metáis takes place on the walls or on stirrers. Deposition of metal on the internal surface of the reactor is undesirable because it causes operating difficulties in collecting the metal. A catalyst may be needed in one médium but not in another. For example, a catalyst is needed for precipitating cobalt and nickel from


159

an ammoniacal sulfate but not from an acid médium. Nickel is precipitated catalytically from an ammoniacal sulfate médium but not from an ammoniacal carbonate médium. Precipitation of a metal may be autocatalytic. Thus, while copper can be precipitated from ammoniacal solution without the need of a catalyst, yet the deposited metal accelerates the process. A difference between the two processes however is that in non-catalyzed reduction, the rate depends on the initial metal ion concentration, while in catalyzed reduction it does not, but depends on the surface área of the catalyst. The commercial precipitation of nickel from the ammonium sulfate system nucleation is induced by adding a small amount of ferrous sulfate which, upon heating to the reaction temperature, hydrolyzes to ferrous hydroxide thus furnishing the catalytic surface required. Nickel deposited in the first step acts as a catalyst for the next. After each reduction, the nickel particles are allowed to settle to the bottom of the autoclave, while the spent solution is drawn off and replaced with fresh pregnant solution. In this way, the nickel particles grow to the desired size, at which point the suspensión is discharged and the nickel powder then separated. There is no need for the ferrous sulfate catalyst in the ammonium carbonate system; as a result, the nickel powder produced in this médium has a lower sulfur and iron impurity level than powder produced from the ammonium sulfate system. Role of additives The presence of certain organic or inorganic substances in the aqueous phase greatly affects the physical nature of metal precipitated. It is possible to precipítate metal powder of certain physical property by simply adding a certain amount of additive. However, when organic additives are used, the carbón content in the powder produced is increased and a special heat treatment is necessary to lower it to 0.01%. Additives may be used for the following purposes:

160


Anti-agglomeration Agglomeration of the precipitated metal partióles may take place, especially at high temperature. This is undesirable because the agglomerated partióles entrap solution causing an impuro produot. Reagents are therefore added to control the partióle size. These are the same as those commonly used to promote uniform growth of cathodes in the electrowinning of metáis, e.g., ammonium polyaorylate, arabic gum, gelatin, dextrin, dextrose, and fatty acids suoh as oleio and steario. These additives are adsorbed on the individual partióles, thus preventing their agglomeration. Smooth surface formation When anthraquinone or its derivativos is added during reduotion, nickel partióles produoed are smooth and regular because of uniform deposition while in the absenoe of anthraquinone, they are coarse and irregular (Figure 6.5). Anthraquinone has no effect on the precipitation of cobalt. The addition of this additive to ammoniacal nickel sulfate or carbonate Solutions also aooelerates the precipitation, and this effect increases with increasing anthraquinone ooncentration up to a certain valué, beyond which it has no further effect.

Figure 6.5 - Effect of anthraquinone on the shape on nickel powder (320 ji); (a) no anthraquinone, (b) in presence of anthraquinone, cross section through smooth nickel sphere after 40 successive depositions

161


,i¿*

-í^.;^-?».

'^í'" 'X '-<'^

l'igure 6.6 - Precipitation of nickel by hydrogen. Left: irregular powder. Middle: course rough powder. Right: Smooth 10 n spheres

Crystalline product Figure 6.7 shows hexagonal platelets of metallic cobalt precipitated in presence of a suitable additive.

Figure 6.7 - Precipitation of cobalt powder by hydrogen

T Pressure Hydrometallurgy

162


163

Industrial application At the Sherritt-Gordon Plant (Figures 6.8 and 6.9), the purified leach solution obtained by ammoniacal pressure leaching of nickel-cobalt sulfide concéntrate contains 45 g/L Ni, 1 g/L Co, 350 g/L ammonium sulfate, and enough free ammonia to give a [NH3] / [NP"" + Co^""] molar ratio of 2. The purified solution is reacted with hydrogen at 3500 kPa and 200°C. Solut on containing 45 g/L Ni anc i g / L Co obtained by ammoniacal teach ng of Ni-Co sulfide ore Nickel precipitation

H2" — - p ,

Figure 6.8 - Foil Saskatchewan near Edmonton, Alberta in Canadá

Y Fiitration Solution containing 1 g/L Ni and 1 g/L Co

^

Nickeí powder

y NiS-CoS precipitation

>

HjS-

y

To ammonium '" sulfate recovery

Fiitration

y

H2SO4 -

Leaching

y íron remo val Fe(OH)3

y Fiitration

y

NHa^

for recycle

.^ Coball oxidation y >•

HJSO4

Nickel ammonium sulfate precipitation

I

y Fiitration

-

y Cobait powder

Niolíel ammonium >- sulfate for nickel recovery

> •

,>

•

1

40

-

\

-

Nickel

conversión

y Hj-

-

•a •g 30

Co^-"••>• Co^* ••

Nickel is precipitated preferentially until its concentration is reduced to about 1 g/L, while all the cobait remains in solution (Figure 6.10). The spent solution containing 1 g/L Ni and 1 g/L Co is then treated with HjS at 80°C and atmospheric pressure, and the precipitated Ni-Co sulfides are filtered off for recovery. The solution is then evaporated to crystallize ammonium sulfate fertilizer. The mixed sulfides precipitated earlier are leached with H2S0^ at 120°C in presence of air at 7200 kPa; acid is used instead of ammonia to avoid the formation of lower oxidation products of sulfur. The solution is purified from traces of iron by adjusting the pH to 5 and filtering off ferric hydroxide.

.2 20

Cobait precipitation

y Fiitration

1 1

Coball ^ powder

y Ammonium sulfate

Figure 6.9 - Recovery of nickel and cobait by precipitation with hydrogen: the Sherritt-Gordon process

=

\

r~^'.

Cobait 1

15

1

V.:--..

30 45 Time, minutes

60

Figure 6.10 - Precipitation of cobait and nickel from ammoniacal solution by hydrogen under pressure

Pre$sure Hydrometallurgy

164

The nickel-cobalt separation is carried out by oxidizing Co(II) to Co(III) to 70°C by air at 700 kPa and in presence of excess ammonia: [Co(NH3),P

[Co(NH JJ3- + e-

YaO^ + Hp + 2e-

20H-

On acidification, nickel ammine complex decomposes and precipitates as the double salt nickel ammonium sulfate:


165

Nickel in Philippines Nickel was recovered from the laterites in Marinduque (Figure 6.11) by roasting, then leaching the calcines in ammonium carbonate according to the Carón process. The basic nickel carbonate cake obtained was dissolved in ammonium sulfate recycle solution and subjected to pressure precipitation by hydrogen in the same way as Sherritt process. Four autoclaves 2.4 x 9.8 m were used. The plant started in 1974 and shut down in 1986. The reason was the energy crisis of the 1970 because all fuel was imported. The company went bankrupt and was taken over by the Government of Philippines ten years later till it was shut down.

[N¡(NH3)J2" + nH^ -^ NP^ + nNH/ while cobalt remains in solution. The slurry is then filtered to recover nickel. The fíltrate, containing cobalt in the cobaltic state, is converted back to the cobaltous state by cobalt powder. This step is essential otherwise a black precipítate of hydrated cobaltic oxide will precipítate during heating. Traces of iron are precipitated by ammonia as ferric hydroxide and separated. Metallic cobalt is then precipitated at 175°C by hydrogen at 2000 kPa in the presence of 25 g/L Co powder as catalyst. After filtration, the solution is then evaporated to crystallize ammonium sulfate. Table 6.1 gives analysis of nickel and cobalt produced by this process.

Pjll Phiiippine Sea

Soutn ^^KyJ < H Chino . ^ ^ B \ . ' ?

Sea WtMÍ r '

LUZON MARtNOUQUE

VAS j J l ^ l ^ 4 jg^^jfiHHM|L' yé^mBBSBj^

'V'

Table 6.1 - Purity of nickel and cobalt produced by hydrogen reduction Cobalt

Nickel Ni Co Cu Fe

99.7-99.85 0.1-0.2 0.01 0.02

Co Ni Cu S

95.7-99.6 0.1-0.5 0-0.02 0.02-0.05

MINDANAO

MALAYSIA

'\ 1

Figure 6 . 1 1 - Marinduque nickel plant in the Philippines

Copper The precipitation of copper from CuSO^ solution takes place through the disproportionation of cuprous ion which has been identified in the course of reaction:


166

2Cu2^ + H2 ^ 2Cu^ + 2H^ 2Cu^ ^ Cu + Cu2^

This leads to low yields oí metal. However, an advantage of this process is that copper can be precipitated from acid solution, i.e., there is no need to add ammonia during precipitation and as a result no ammonium sulfate is produced as the case with nickel and cobalt (Figure 6.12). 40 \

\ \

• 413 K

*

• 423 K

35 •

\^*

167

Copper and zinc A flotation concéntrate of copper-zinc-iron sulfide is leached with ammonia at 90°C under an air pressure of 700 kPa. Copper and zinc pass into solution as ammines, while iron is precipitated as hydrated ferric oxide and filtered. Sulfamates formed during leaching are oxidized and hydrolyzed to sulfate at 230°C and 3500 kPa with air. The molar ratio of free ammonia to copper is decreased to Vi by adding sulfuric acid to the autoclave prior to reduction by hydrogen to precipítate copper. Small amounts of ammonium polyacrylate are added to permit control of the physical characteristics of the powder produced. The solution is then treated with carbón dioxide under 700 kPa at 37°C to precipítate zinc hydroxy carbonate, which is then ñltered off:

A433K

30

"&


• 443K

\ •

-

2Zn(NH3)2SO, + CO^ + SHÔ ^ Zn(OH)2.ZnC03 + 2(NH,)2SO,

••

= 25

"o

c .2 SOIS

"S

•

The clarified solution is evaporated to crystallize ammonium sulfate which is marketed as a fertilizer.

*«,^^^

•

•>

'—¡

0

0

"'X~^

10 5

40

60

80

100

140

160

180

Time, min

Figure 6.12 - Precipitation of copper by hydrogen under pressure in the range 140-170°C

Copper scrap or cement copper is dissolved either in ammoniacal ammonium carbonate at 60°C at atmospheric pressure with continuous aeration, or in dilute sulfuric acid. When ammoniacal médium is used, the molar ratio [NH3]/ [Cu^-^] should be equal to 2.4. After filtration to remove insoluble material, a small amount of anti-agglomerating agent is added, then solution is heated to 200°C under 6000 kPa hydrogen. The copper powder precipitated is filtered off, washed, and then dried in a reducing atmosphere at 600°C.

Precipitation of metáis from non-aqueous médium Many metal ions are extracted by organic solvents by forming a coordination bond. When this loaded organic phase is treated by hydrogen at high temperature and pressure in an autoclave, the metal precipitates in powder form and the organic phase is regenerated. The process may be described as precipitation by substitution since no ionic species are taking part in the reaction as compared to precipitation by hydrogen from an aqueous phase. The substitution reaction can be represented as follows: H2(g) R,M, 2

(org)

+K, , 2(org)

H2(org) 2RH, (org)

+M,, (s)

where RH is the organic solvent and M is a divalent metal. A typical example is the precipitation of metallic copper powder from hydroxy-quinoline-kerosene phase containing copper:


168

+H

+ Cu(s)

-^2

2(g)

(org) (org)

Uranium oxide from leach solution


169

lets and each tower is operated continuously until 10 tons of product has accumulated. The reduction end solution which contains only 3 to 5 mg/L uranium is recycled to the pressure leaching state. Precipitation by carbón monoxide Carbón monoxide has been used for precipitating silver from AgNO solution and copper from [Cu(NH3)J^"' solutions obtained by leaching brass scrap in ammoniacal ammonium carbonate : [Cu(NH J / - -^ Cu2- + 4NH,

The hydrothermal leaching of certain uranium ones with sodium carbonate:

Cu2^ + CO + H p ^ Cu + CO2 + 2H" UO2 -^ UOj^" + 2eUO/^ + 3CO32- -. [UOjíCOg)/YzO^ + HjO + 2e- -^ 20H-

Precipitation takes place at 150°C with CO partial pressure of 5,000 kPa. The solution is then boiled to precipítate zinc as basic carbonate. Precipitation of metáis by CO is much slower than by hydrogen This may be due to the fact that CO first reacts with water to form hydrogen:

Uranium dioxide is recovered from the uranyl carbonate leach solution by precipitation with hydrogen under pressure: [UO^ÍCOg)^^ -^ UO/^ + 3CO32UO/^ + 2e- ^ UO2 H2 -^ 2W + 2e-

CO + up ^ H^ + CO2 Precipitation by sulfur dioxide On passing SO^ into a solution of copper sulfate at room temperatura copper sulfite will precipítate. However, if precipitation is carried out at 150°C and 350 kPa, metallic copper is precipitated according lo: ^

Overall reaction: C u s o , + SO, + 2H O ^ Cu + 2H SO. [UO^ÍCOg)^^ + H2 -^ UO2 + 2HCO3- + CO32-

or At Kalna in former Yugoslavia, the reaction is conducted at 150°C and 1500 kPa in vertical autoclaves containing pellets of partly sintered UO, as catalyst. The precipítate builds up on the catalyst pel-

Cu2^ + SO.2- + H,0 - ^ Cu + 2W + SO


170

The drawback to this process is the low yield of copper as shown in Figure 6.13. The corrosión problems due to the acidic environment, and the presence of small amounts of sulfur in the copper produced. The low yield may be due to the intermedíate formation of cuprous ion which disproportionates precipitating half of the copper and regenerating the other half as cupric ion: Cu2^ + e- - ^ Cu* 2Cu* ^ Cu + Cu2*

PL

%apter 6 - Precipitation

171

.is formed. The following reaction takes place: Cu2* + SO 2- + 2NH, + KO -^ Cu + 2 N H / + SO 23

3

2

4

4

This resulted in complete precipitation of copper and has the advanlage of operating under basic conditions thus eliminating corrosión problems. However, it has the inconvenience of producing ammoiiium sulfate which has to be marketed as fértilizer. In a similar way, cuprous ion in a copper ammine sulfite solution can be reduced to metallic copper when the pH is adjusted to 3 by sulfurous acid to precipítate the double salt CU2S03.(NH^)2S03 which, when slurried with water and heated at 150°C in an autoclave:

100

Cu" + e- - ^ Cu

SO32- + Hp -^ SO/- + 2H" + 2eSO32- + 2H* -^ SO2 + Hp Overall reaction: 2 3 Time, hours

4

Figure 6.13 - Precipitation of metallic copper from CuS04 solution by SO^

The decomposition of sulfurous acid at the reaction temperature to HjSO^ and elemental sulfur according to the equation: 3H2SO3 -^ 2H2SO, + S + HÔ

accounts for the presence of small amounts of sulfur in the copper produced. This process was improved by adding an ammoniacal solution of ammonium sulfite instead of SOj, i.e., neutralizing the acid as soon

Cu,S03.(NHJ,SO

2Cu + SO, + 2 N H / + S O / 2

4

4

Sulfur dioxide generated must be vented during heating and collected for recycling. The process, however, was developed up to the pilot stage only at Anaconda Research in Tucson, Arizona.

lONIC PRECIPITATION Precipitation by hydrogen sulfide Hydrogen sulfide is a poisonous and corrosive gas. In certain concentrations, it explodes in air. In precipitating metal sulñdes by H2S, the following points should be taken into consideration.


172

Acid generation. Acid is generated during precipitation: M2^ + H^S ^ MS + 2H" and can be used in the leaching circuit, or must be neutralized before disposal. Polymorphic precipitates. Cobalt and nickel sulfides exist in several polymorphic forms with different solubilities. The alpha forms, obtained by precipitation from basic solutions, are amorphous and soluble in dilute acids. The beta forms, precipitated from weakly acid solutions, are crystalline and only slightly soluble in 0.1 M HCI.


173

Catalysis. Precipitation may be greatly accelerated by the addition of a catalyst. Thus, the precipitation of nickel sulfide from weakly acidic solution is extremely slow, but the addition of small amounts of iron or nickel powder accelerates the reaction greatly especially at high temperature and pressure. A process developed at Moa Plant in Cuba for recovering nickel and cobalt was based on this principie. The ore containing 1.35 % Ni and 0.15 % Co is leached with sulfuric acid, filtered, and the solution treated with H^S at 120°C and 1000 kPa in the presence of iron powder as catalyst to precipítate NiS and CoS. Precipitation is conducted in pressure reactors. The mixed sulfide is then filtered, dried, and shipped for further processing.

PRECIPITATION OF IRON OXIDE NAKE-UP HjS FHtiH Hj-HgS PLANT

LIMOR FROM KEUTRAUZATiaít LIQUOR PREHEATER (LOCATEB IK LEACHING ASEA)

Ferric ion is present in many solutions and there is always interest to remove this iron. Ferric ion hydrolyses according to: Fe^^ + Hp ^ Fe(0H)2^ + H^ Fe(0H)2^ + Hp ^ Fe(OH)* + H^ Fe(OH)* + HjO ^ FeOOH + H^

» PS!6 STEAN

At high temperature FeOOH is transformed into Fe203: 2FeOOH ^ Fe203 + Hp

OFF H j S COOLER

ÁREA PifODUCT STMAGE

yL.

BARÜEM Ll TO WASTE PfiflOUCT SULPHIDE Bt TRUCK TO PORT ÁREA ST08ASE

Figure 6.14 - Precipitation of nickel and cobalt sulfides by H^S

In chloride media, 3-FeOOH is produced in preference to goethite (a-FeOOH) produced in sulfate mdium. It has excellent filtration properties.

174


Jarosite Sometimes a hydroxyl iron salt is formed similar to the naturally occurring mineral jarosite. The ñame is derived from the locality Barranco del Jaroso in the Sierra Almagrera in Almería, Spain where mineral was first found and described. These compounds are crystalline and easy to filter and wash. For example, ferrihydroxy sulfate precipitates according to: 3Fe3^ + 3S0/- + S H p


175

the ferrihydroxy sulfate. Figure 6.15 shows the conditions for the precipitation of hydroxy salts of iron. Jarosite can be converted to hematite by heating in an autoclave at 220°C.

Fe(OH)3.Fe,(SOj3 + 3H^

This precipítate can also be represented as: Fe2O3.2SO3.H2O or Fe(OH)SO^. There are numerous hydroxy ferric sulfates that form under different conditions and can be represented by the formula Fe2O3.xSO3.yH2O as shown in Table 6.2. Table 6.2 - Stable solid compoimds in the system FeÔj-SOj-RjO at Fe-'* ion and HjSO^ concentrations below 100 g/L in the temperature ranga 75-200°C Solid phase FezOaxSOsyHjO

Formula

Ñame

X

y

0

0

FejOa

Fe203

Hematlte

0

1

FeaOjHsO

FeOOH

Goethite

1/2

5/2

FesOaVzSOa-^HsO

Fe4S04(OH)io

Glockerite

4/3

3

Fe203''/3S03-3H20

(H30)Fe3(S04)2(OH)6

Hydronium jarosite

1-

Fe203-2S03H20

Fe(OH)3Fe2(S04)3 or F e S 0 4 ( 0 H )

Ferrihydroxy sulfate

At the special composition when x = 4/3 and y = 3, the compound formed is known as hydronium jarosite. In this compound, monovalent ions such as Na\ K^ NH^ôr Ag'^mayreplacethehydrogenwhile divalent ions such as Pb^* and Hg^^ may replace the iron as shown in the Table. Silver even when present in very low concentrations is selectively incorporated into the jarosite precipítate. Under such circumstances, the formation of jarosite may be a nuisance because it represents a loss of the non-ferrous metal and a contamination of

Figure 6.15 - Conditions for the precipitation of iron oxide, oxide hydroxide, hydroxide, and hydroxy salts from 0.5M ferric sulfate solution

Chloride system Contrary to the sulfate system, silver and lead are not precipitated in the jarosites formed in chloride media, even when present in high concentrations. This is likely owing to the presence of chloro-complexes that have the large size and charge to be incorporated into the jarosite structure. Akita Zinc In Akita Zinc in Japan (Figures 6.16 and 6.17) the zinc concéntrate is roasted, leached, and zinc is recovered by electrowinning. The residue from leaching mainly zinc ferrite and gangue is heated in autoclaves to 115°C in sulfuric acid and SO2 so that iron will be present in the ferrous state. This treatment is necessary so that after filtration to remove the gangue, gallium and indium are recovered from solution. The solution is then heated at 200°C in autoclaves in presence of oxygen Figure 6.16 to precipítate Fe203. Zinc is then recovered Location of Akita from solution (Figures 6.18).

TT 176


177


Ruhr-Zink Ruhr-Zink GmbH (Figures 6.19 and 6.20) a subsidiary of Metallgesellschaft was founded in 1968 in Datteln. The neutral leach residue of zinc calcine is leached in H.SO^ with concomitant additions of 2

concéntrate to decompose the ferrites and to reduce all the iron to the Fe^* state. Sulfate and some excess acid are controlled by the addition of lime to precipítate gypsum. Finally, the ferrous sulfate solution is heated in an autoclave to around 200°C with oxygen injection to precipítate Fe203. The solution must be reduced before heating to avoid the formation of basic iron sulfates in the autoclave. Pressure leaching commenced operation at the Ruhr-Zink refinery in 1991 integrated with the existing roast-leach-electrowinning. It closed in 2008 due to the collapse in the zinc price, as well as Germany's very high electricity prices.

Figure 6.17- Akita Zinc in Japan Zinc calcine leacli residue Spent electrolyte —

— V

V

SO,

V

Pressure leaching

+ S/L Separation Cao,

4

1 \ pH 1.5

\' Gypsum - ^ -

S/L Separation

CU^*

1 \ r

- • " Pb-Ag residue Zn

Cementation

'' S/L Separation

- * - CUjAs

CaCOj

-.

\ pH4.4

1 S/L Separation O,

^'

-*^ Ga-ln hydroxides

\'

Precipitation

+ S/L Separation

->~ ZnSO, solution 4

Fe

k

Figure 6.18- Precipitation of FeÔj in zinc industry in Japan by oxygen in autoclave

Figure 6.19 - Location of Datteln north of Dortmund in the Ruhr District of Germany


775


179

10000 £000

1000

•

500

, 100

Chulchlanl*e¥ Tozcwa «t. «L Makhmeiov et •!. Robini A Qlcilrat Krkuie & Ettel Dov» & Rímctidt KniUM & gttal

(1956) (1«78) (1981) (1987) (1987) (1985) (1987)

AMORPHOUS (7) FeAi04

O) SO

-1

-2

-a

b

tf.

_1 ffl

=> _l 0 co

/ / •

10

o

5 Natural Sunpl*

«e

< Figure 6.20 - Ruhr-Zink

PRECIPITATION OF ARSENATE

<

1.0

0.5 CRYSTALLINE SCOflOOlTE

0,1 0.05

The disposal of arsenic has been accomplished in practice by the formation of metal arsenates because of their low solubility. For example, calcium arsenate has been precipitated by adding lime to the solution:

3CaO + 6H" + 2AsO/- CaJAsOJ, + 3HP However, under the influence of CO2 in the air, calcium arsenate decomposes to calcium carbonate liberating arsenic oxide that goes into solution. That is why the formation of more stable ferric arsenate known as scorodite (FeAs04.2H20) has been adopted. A minimum molar Fe:As of 4 is required in the solution to favour the formation of an amorphous scorodite (Figure 6.21).

H,|A«04

HjAtO;

,HAtO/

4

PH Figure 6.21 - Solubility of scorodite formed at atmospheric [upper curve] and at high temperature [lower curve] as a function of pH

Crystalline scorodite is less soluble than amorphous ferric arsenate. Precipitation in atmospheric pressure at 95°C with addition of a seed, yields about 90% of arsenic for solutions containing between 5 to 10 g/L As(V) as scorodite. However, at 150-190°C complete arsenic precipitation for solution containing 5 g/L As(V) takes place, Fe/As molar ratio: 1:1 to 9:1. Crystalline ferric arsenate has lower solubility than the amorphous ferric arsenate.

180


Stability and solubility ofCaJAsO^)^ The solubihty of Ca3(As04)2 decreased one to two orders of magnitude over the range of pH 9-12.6 and in the presence of phosphate. However, industrial application of this compound has not been used in industry.

Attempts to Avoid Autoclaves

Arsenic disposal from copper smelter dusts Dust is leached with spent electrolyte then after recovery of Zn, Pb, Cd, Ag, arsenic is disposed of in the form of a ferric arsenate/gypsum sludge. Uses Arsenic trioxide, 99.9% Asp^ is the main commercial arsenic compound recovered by roasting of copper or cobaU concentrates and leaching of copper smelter dusts. It is principally used for the production of wood preservatives.

Introduction The Use of Bacteria in Leaching Concentrates Application to chalcopyrite concentrates Geocoat process Atmospheric Leaching Using Different Reactors Treadwell process for chalcopyrite Arbiter process HydroCopper process Galvanox process Cuprex process Zinc concentrates Albion process for gold

181 182 182 184 186 186 188 190 191 192 193 194

INTRODUCTION There have been many attempts to avoid the use of autoclaves in the erroneous belief that these equipment are expensive, difficult to opérate and maintain, are dangerous, and cannot compete with standard equipment because of their small capacity. In addition, it was argued that pressure leaching of nickel is possible because nickel is an expensive metal but it would not be possible to use this technology for a cheap metal like copper. The increase in size, in number, and success of plants using autoclaves not only for leaching concentrates but also for leaching low-grade ores attest to the faulty misconception of pressure hydrometallurgy.


182

If the price of nickel is five times the price of copper then it should be possible to use autoclaves for leaching copper concentrates containing 5% Cu - but copper concentrates containing 20% Cu is already available. Two approaches, however, took place to avoid the use of autoclaves: the use of bacteria in leaching concentrates and the adoption of atmospheric leaching using different reactors and different conditions.

Chapter 7 - Attempts to Avoid Autoclaves

183

Peñoles in Monterrey in México (Figure 7.1). The plant operated for a year with a capacity of 200 tpa copper cathode production using commercial equipment, and demonstrated well the technical feasibility of a totally integrated process with high levéis of copper recovery from arsenic-containing concéntrate blend. However, due to the low price of copper at that time BacTech did not go forward with the project to a commercial scale.

THE USE OF BACTERIA IN LEACHING CONCENTRATES It was believed that bacteria used in aiding the leaching of low-grade ores can as well be used to leach concentrates especially chalcopyrite concentrates. Bacterial leaching has been successfully applied for heap leaching of low-grade copper ores. One of the major operations is that at Bingham Canyon in Utah. The leach solution coUected at the bottom of the heap is extracted by organic solvents and the strip solution is electrolyzed to get high purity copper. The process was extended to treat auriferous pyrite concentrates to libérate gold and render it amenable to cyanidation by a process known as BIOX. Application fo chalcopyrite concentrates In the past few years there has been interest to apply bacterial leaching to treat chalcopyrite concentrates. For example, a continuous small scale bioleach pilot plant was established in 1998 by BacTech at Mt. Lyell in Tasmania. The plant was integrated with solvent extraction and electrowinning for copper recovery. Peñoles plant A demonstration scale plant was constructed in 2001 by the joint technology partnership of BacTech and Mintek in conjunction with

Figure 7.1 - Pilot plant for bioleaching of chalcopyrite concentrates Peñoles in Monterrey in México

Alliance Copper plant In 2002, Alliance Copper, which is a joint venture between BHP Billiton and Codelco in Chile, also built a 20,000 tonnes/year demonstration plant near Chuquicamata (30 km from Calama) in Chile for $50 million (figure 7.2). The plant was composed of six large reactors, mechanically agitated, and lined with acid-resisting brick. Since a thermophilic bacteria was used in the system it was possible to opérate at a temperature of about 80°C and this accelerated the reaction. Since oxygen was used and not air, it was necessary to introduce COj in the tanks because bacteria require this gas to build up the cell structure. Also, it was necessary to add phosphates and ammonium ion needed by the bacteria as nutrients. The reaction


184

was slow; it was complete in 4-5 days. The plant was shut down few months later.

Chapter 7 - Attempts to AvoidAutoclaves

CuFeS^ + 4O2

185

CuSO, + FeSO, 4

4

From this it can be seen that a large amount of oxygen will be consumed, a large amount of lime will be needed to precipítate ferrous sulfate, and there will be an excessive disposal and material handling problem of ferrous hydroxide-gypsum mixture: FeSO^ + Ca(0H)2 + 2 H p -> FeíOH)^ + CaS0^.2H20 In the recovery step by electrolysis, acid will be generated and must be disposed of Actually, all attempts to apply this technology for chalcopyrite concentrates have failed. When bioleaching technology is compared with pressure leaching, the reaction that takes place in one autoclave at 150°C and 1000 kPa oxygen partial pressure is as follows: 2CuFeS2 + 4H* + 2720^

Figure 7.2 - AUiance Copper pilot plant in Chuquicamata

2Cu2* + Fe,0, + 4S + 2H.0 •'2"'3

The advantages of this route are the following:

Geocoat process The Geocoat process was developed by GeoBiotics in Lakewood, Colorado. In this process, copper sulfide flotation concéntrate slurry is coated onto a crushed and sized carrier rock which may be barren or may contain sulfide or oxide mineral valúes. The coated material is stacked on an impervious pad for biooxidation. High temperature thermophile miicroorganisms are used to accelerate copper leaching. It does not make sense however that after concentrating an ore the concéntrate obtained is then diluted by mixing with a crushed rock for heap leaching. The concéntrate should be suitable for agitation leaching. Drawback of bacterial leaching plañís In spite of this enthusiasm for bioleaching technology, one cannot recommend its use for leaching chalcopyrite concentrates because it cannot be economical for the foUowing reasons. The leaching reaction for chalcopyrite is as follows:

• The reaction is fast, complete in 20-30 minutes • Oxygen consumption is VA moles per mole chalcopyrite as compared to 4 moles in the case of bioleaching, that is less than one third that required for bacterial leaching • One reactor is enough • Cu^"" is already separated from Fe^"' since FejOg is precipitated during the reaction • All the sulfur in the concéntrate can be obtained in the elemental form • When copper is recovered from solution by electrowinning, the acid generated at the anode is equal to that required for leaching henee no acid disposal problem • There is no material handling and disposal problem involving lime addition • Any arsenic present in the concéntrate will remain in the residue as ferric arsenate

T'

186


'TT

187


ATMOSPHERIC LEACHING USING DIFFERENT REACTORS

In this process (Figure 7.4), chalcopyrite concéntrate was treated with concentrated sulfuric acid at 200°C:

When leaching concentrates at ambient pressure and at high temperature but below the boiling point, then it is necessary to use a reflux condenser to avoid the escape of vapours. If the reacting mass solidifies after few minutes when the reaction started as for example in the case of chalcopyrite, then the baking process should be adopted and special equipment must be used.

CuFeS, + 4H,S0^ -^ CuSO^ + FeSO, + 2 S 0 , + 2S + 4 H , 0 2

2

4

4

4

2

2

Sulfide concéntrate Make up HzSO^

Concentrated HsSO^

V Y t , so. , 7 " Baking H20-

y

[-

-"^

Acid plant

I |

V Leaching

Treadwell process for chalcopyrite The action of concentrated H2S0^ on sulfide minerals received attention in the 1960s because of the fact that under certain conditions elemental sulfur can be formed and therefore pollution due to SOj that generally forms in smelters can be avoided. A process was developed on laboratory scale at Treadwell Corporation in Bronx, New York and was tested in Tucson, Arizona in 1970 at Anaconda Company. A 100 tonnes chalcopyrite concéntrate per day pilot plant was constructed for this purpose (Figure 7.3).

Filtralion

[

— • - S u l f u r , gangue

Purification

I

Spsnt electtí)lyte Evapofatlon \^— --|

I Recovery

|

Y Metal

Figure 7.4 - Flowsheet for the treatment of chalcopyrite with concentrated sulfuric acid

Copper and iron in the mineral are converted to water-soluble sulfates, while elemental sulfur is formed. The sulfur dioxide formed during the reaction must be converted to sulfuric acid, for recycle. A number of side reactions may also take place if the conditions are not properly selected. By using a two-stage process whereby the concéntrate was first agitated with a stoichiometric volume of 98% sulfuric acid. This step was quite short, taking only few minutes. In the second stage, the solidified mass was heated further until the reaction was complete and the product then leached with water to remove the soluble sulfates from the gangue and elemental sulfur. A complicated system of bucket elevator and silica balls for heat transfer was used. The process was criticized by many metallurgists because SO2 was generated and must be converted to acid for recycle and ferrous sulfate must be decomposed to genérate acid for recycle.

Figure 7.3 - Pilot plant at Tucson, Arizona

188


The situation was compounded further by the poHtical situation in Chile where the new socialist regime nationalized the copper industry and Anaconda lost its properties in Chuquicamata. The pilot plant was abandoned, and the process was dismissed as uneconomical. At the same time, a new process was developed and became known later as the Arbiter process.

189


is based on leaching chalcopyrite concéntrales with ammonia at 75-80°C in presence of oxygen to form copper ammine sulfate and i ron hydroxide (Figure 7.6): 2CuFeS2 + I2NH3 + SyzOj + 2 H p 2[Cu(NH3)J2* + 4 N H / + 4S0/- + ?ep^

Incidentally, researchers at Kennecott Copper Corporation in Salt Lake City, Utah had a similar process under investigation but they used a bug mili for treating the sulfuric acid-chalcopyrite concéntrate and the process was abandoned before leaching a pilot scale.

Chalcopyrite concéntrate

NH,

Air

^ 2 ' ^ Leaching

"1~ Eievation — __

fü

Solids

Filtration

CaO

ir

V

Distillation

Raffinate

M—Water

Washing

Solvent Extraction

. Dilute NH3 solution

Filtration

Organic Residue

Gypsum Water-

Washing

Stripping

yfp

Plan

Figure 7.5 - Bug mili

Arbiter process The process (Figure 7.X) was developed in 1970 in a pilot plant at Anaconda Company in Tucson under the direction of Nathaniel Arbiter, a former professor of mineral processing at Columbia School of Mines in New York City. A commercial 90 tonnes/day copper plant went in operation at Anaconda, Montana few months later. The leaching plant was composed of 10 intensely agitated vessels 14 m3 each and 5 counter-current decantation thickeners, with the first overflow filtered to form the pregnant solution. The process

Organic phase H,SO,

CuSO^ solution Electrowinning

Copper

Figure 7.6 - Arbiter process

After filtering the solids, copper was extracted by LIX-65N, stripped by sulfuric acid, and electrolyzed to give metallic copper. Lime is then added to the raffinate, and the slurry boiled to distil off ammonia for recovery and precipítate gypsum for disposal: (NHASO + Ca(OHL

2NH, + CaS0,.2H,0 3

4

2

TT 190


This ammonia recovery step as well as that from the residue washings proved to be technically difficult and economically unsound and was the main reason for the shut down. HydroCopper process Chemists at Outotec, formerly Outokumpu Research Oy, in Pori, Finland have developed HydroCopper process for the treatment of copper sulfide concentrates avoiding the use of autoclaves (Figure 7.7). The process is based on leaching the concentrates in a strong NaCl solution containing Cu^* ion at pH 1.5-2.5 in agitated reactor at 85-95°C in presence of oxygen. Copper goes into solution as Cu* while iron is precipitated as hydroxide. After filtration and solution purification NaOH is added to precipítate CU2O, which is then slurried in water and reduced in autoclaves by hydrogen under pressure. Copper sulfide concéntrate HCI + NaCl

'•

'

r

Leaching, 90°

'' S/L Separation

' Impurities

-^—

^

r^

sidue

Cl ^'2

Leaching of sulfides in chioride media has been tried before in Clear Process developed by Duval Corporation in Arizona in the 1970s. The process, however, was not industrialized because the electrowinning of copper from chioride médium produced dendridic powder contaminated by silver that was difficult to process further. It seems that this was the reason for Outokumpu chemists to avoid the electrowinning route and consider the production of CuÔ and its reduction. The precipitation of copper by hydrogen under pressure from aqueous chioride system is not effective. Another option is the thermal reduction of solid CuCI in a fluidized bed. Outokumpu researchers also found out that CUjO disproportionate in dilute H2SO4: Cup + 2H*

Cu + Cu"* + Hp

CuSO^ formed can be reduced by hydrogen in an autoclave by known methods. In this process, CaCI^ was a waste product for disposal. In the Outokumpu process NaOH is used instead of Ca(0H)2 so that NaCl produced can be electrolyzed to recover hydrogen for reduction, NaOH for precipitation, and chlorine transformed into HCI. Iñ other words, an important section of the process is the regeneration of the reagents. Galvanox process

r

'' Precipitation

1

NaCl solution

Filtration

Hp—.-

191

, Solution

Purification

\


,

NaOH

Electrolysis

Cup

Slurrying y'

'

H?

Reduction

T Copper

Figure 7.7 - Outokumpu, now Outotec, HydroCopper process

In the Galvanox process leaching of copper concentrates is conducted at 80°C under atmospheric conditions using ferric sulfate as oxidant in presence of pyrite. The process takes advantage of the galvanic couple between pyrite and chalcopyrite which accelerates the rate of leaching (Figure 7.8). Elemental sulfur is produced. However, to achieve this accelerating effect, large amounts of pyrite must be added which involves material handling problems and increased size of reactors.

192 Pressure Hydrometallur^

Chalcopyrite


193

Pyrjte

4Fe^-

4Fe=-

Figure 7.8 - The galvanic couple between pyrite and chalcop3TÍte Galvanox process

by adding sodium chloride, which enhances conductivity and minimizes the possibility of the precipitation of cuprous chloride, and the solution is sent to the Metchlor electrolysis cell. The cell has two compartments separated by a catión exchange membrane. It uses a titanium cathode and an inert anode. Copper granules are deposited in the cathode compartment, and the electronic balance in the catholyte is maintained by transfer of sodium ions from the anolyte through the membrane.

Cuprex process The Cuprex process was developed in 1988 as a joint venture between Técnicas Reunidas in Spain, ICI in United Kingdom, and Nerco Minarais. It uses ICI's selective DS 5443 extractant and Técnicas Reunidas' Metchlor electrowinning cell. Severa! two-to three-week runs on a variety of copper concentrates were completed at Técnicas Reunidas's facilities in Madrid, and the process was considered ready for commercialization. The advantage of this new extractant is that it permits the recovery of copper from chloride médium without the necessity to transfer from chloride to sulfate. In this process, copper concéntrate is leached in two stages with ferric chloride solution at 95°C and atmospheric pressure to produce a solution of cupric chloride, ferrous chloride, and elemental sulfur: CuFeS^ + 4Fe3* -> Cu^^ + SFe^* + 2S The leach residuo consists of gangue, pyrite and sulfur. The pregnant solution, is sent to the solvent extraction circuit where it is contacted with a kerosene solution of DS 5443. Using three extraction stages, copper in the aqueous phase is reduced to less than 0.5 g/L. The loaded organic is scrubhed with spent anolyte from the electrowinning cell and then stripped by contacting with water at 65°C. Three strip stages produce an aqueous copper chloride solution grading over 90 g/L Cu. The chloride ion content is then increased

Spent catholyte leaving the cell is a sodium chloride solution containing about 10 g/L Cu roughly divided between the cupric and cuprous oxidation states. This proceeds to the reforming stage where it is treated with chlorine gas from the anode compartment of the electrowinning cell to oxidize the cuprous ions to cupric. Cupric chloride is removed from the reformed catholyte in the depletion section by contacting it with copper-free organic from the stripping section of the solvent extraction. A two-stage depletion at a high 8:1 organic/aqueous ratio reduces copper in the spent catholyte to less than 0.1 g/L. The organic, containing relatively little copper, is recycled to the extraction unit and the aqueous raffinate becomes anolyte in the electrowinning cell. Any silver in the original concéntrate reports in the raffinate from the solvent extraction stage and can be recovered by zinc dust precipitation. Excess iron from the leaching of chalcopyrite is removed as goethite in a subsequent pressure oxidation step which simultaneously regenerates part of the leachant. The remaining ferric chloride leachant is regenerated by chlorine from the electrowinning cell. Zinc concentrates In spite of the success of the aqueous oxidation process of zinc sulfide in three operating plants, there are still attempts to avoid using autoclaves. For example, a plant under construction in San Luis Potosi in México will use a Finnish technology that uses four large

194

Chapter 7 - Attempts to AvoidAutoclaves


195

jet. Special fine grinding equipment is used (Figure 7.10). As an approximate guide, it is expected 2-2.5% oxidation of the sulfide matrix per hour. To achieve 80% oxidation requires 34-40 hours for a typical refractory gold sulfide concéntrate at a grind size of 80% passing 11 microns. This compares with 2-3 hours in an autoclave for a material ground to 80%) passing 44 microns.

reactors operating at 90°C instead of one autoclave (Figure 7.9). Instead of two hours residence time in an autoclave the múltiple reactors will be operating for 12 hours under continuous air injection. The reaction taking place: ZnS + 'ÁO^ + 2H^ ^ Zn2- + S + H p

It is believed that such operation cannot compete with pressure leaching. OFF

ON

. Ah

Gas hold-up

Measurement iocation

Figure 7.10 -Albion fine grinding mili

Motor

Gas inlet

Motor I

The process was originally designed for zinc sulfide and extended to refractory gold ores. It is claimed that the capital cost of an Albion plant can be lower than a comparable pressure or bacterial leach, due to the simplicity of the process flowsheet. It should be noted however that the solubility of oxygen at 90°C at atmospheric pressure is low and no data are known regarding the filtration of the fine residue.

Gas inlet

Figure 7.9 -A reactor designed for leaching zinc sulfide concéntrate at 90°C. Leñ: before introducing oxygen, right: after introducing oxygen

Albion process for gold To avoid using autoclaves researchers proposed the Albion process for liberating gold from pyrite. The process is named after the suburb where it was developed. It involves fine grinding of ore and using oxygen in leaching at atmospheric pressure in conventional agitated tanks at 90°C. Oxygen is introduced by a special supersonic gas

>

196


8 Laboratory Autoclaves and Pilot Plants Introduction Laboratory Autoclaves Materials of Agitated autoclaves Zipper Clave High torque magnetic drives Acid digestión bombs Engineering Aspects Pilot Plants

construction

197 198 198 202 213 218 220 221 225

INTRODUCTION Laboratory autoclaves for hydrothermal investigations are available in a variety of sizes, models, and materials of constructions. They vary in sizes from 25 mL to 2 L for laboratory studies and 5 to 50 gallons for pilot plant work. They are essential tools for studying aqueous oxidation of sulfide concentrates, dissolution of oxide minerals at high temperature and pressure and hydrothermal precipitation reactions. The máximum pressure and temperature at which any pressure vessel can be used will depend upon the design of the vessel and the materials used in its construction. Since all materials lose strength at elevated temperatures, any pressure rating must be stated in terms of the temperature at which it applies. A review of existing models and their accessories will be given.

198


LABORATORY AUTOCLAVES

Table 8.1 - Materials of construction for Parr laboratory autoclaves Major elements, %

Materials of construction The choice of the material of construction (Table 8.1) of an autoclave depends on the operating médium whether acidic or alkaline, the temperature range, and the presence or absence of oxidizing atmosphere. Table 8.2 provides a set of multipliers which can be used to convert 350°C pressure ratings for any T316SS vessel to higher or lower temperatures. It can also be used to determine the corresponding ratings for vessels of the same design made of other materials. No reactor or pressure vessel should be operated above these máximum temperature limits.

Fe

Ni

Cr

Mo

Mn

Other

T316 Stainless Steel

65

12

17

25

2

Si 1.0

Carpenter 20Cb3

35

34

20

25

2

Monel 400

1.2

66

Inconel 600

8

76

15.5

HastelloyB -2

2

66

1

28

6.5

15.5

16

1

HastelloyC -276 Nickel 200 Titanium Grade 4 ZIrconlum Grade 705

Stainless Steel 316 At ambient temperatures stainless steel 316 offers useful resistance to dilute sulfuric, sulfurous, phosphoric and nitric acids, but sulfuric, phosphoric and nitric acids readily attack T316SS at elevated temperatures and pressures. Halogen acids attack all forms of stainless steel rapidly, even at low temperatures and in dilute solutions. Although T316SS offers excellent resistance to surface corrosión by caustic stress corrosión cracking can occur in pressure vessels. This phenomenon begins to appear at temperatures just above 100°C. T316SS offer good resistance to ammonia and to most ammonia compounds, Halogen salts can cause severe pitting in all stainless steels. Chlorides can cause stress corrosión cracking, but many other salt solutions can be handled in stainless vessels, particularly neutral or alkaline salts. Carpenter 20Cb-3 is an enriched grade of stainless steel, designed specifically for use with dilute (up to 30%) sulfuric acid at elevated temperatures. It can also be used for nitric and phosphoric acid systems as well as for all Systems for which T316SS is suitable.

199

Chapter 8 - Lab Autoclaves & Pilot Plants

Cu 3.5, Nbl.O max. Cu 31.5

1

Co 1 W4.0, Co 2.5

99 Commercially puré titanium

Ti 99 min.

Zr 95.5 min, Hf4.5 max,Co2.5

Table 8.2 - Pressure rating rectors (Parr Instrument Company) Temperature, °C 25

100

200

300

350

T316 Stainless Steel

1.13

1.13

1.09

1.04

1.00

Monet 400

1.20

1.20

1.20

1.20

1.19

Inconel 600

1.20

1.20

1.20

1.20

1.20

Hastelloy B - 2

1.20

1.20

1.20

1.20

1.20

HastelloyC -276

1.20

1.20

1.20

1.20

1.20

Nickel 200

0.60

0.60

0.60

0.60

Carpenter 20Cb3

1.20

1.20

1.17

1.16

Titanium Grade 2

0.75

0.64

0.51

0.36

0.34at316°C

Titanium Grade 4

1.20

1.20

0.81

0.63

O.eOat 16°C

Zirconium Grade 705

1.20

0.98

0.76

0.65

0.61

0.60at316°C 1.16

200


Monel 400 is an alloy comprised essentially of two-thirds nickel and one third copper. For many applications it offers about the same corrosión resistance as nickel, but with higher máximum working pressures and temperatures and at a lower cost because of its greatly improved machinability. It is widely used for caustic solutions because it is not subject of stress corrosión cracking in most applications including the pressure of chloride salts. It is also an excellent material for fluorine, hydrogen fluoride and hydrofluoric acid systems. It offers some resistance to hydrochloric and sulfuric acids at modest temperatures and concentrations, but it is seldom the material of cholee for these acids. As would be expected from its high copper content, Monel 400 is rapidly attacked by nitric acid and ammonia systems. Inconel 600 is a high nickel alloy offering excellent resistance to caustic and chlorides at high temperatures and high pressures when sulfur compounds are present. In caustic environments, Inconel 600 is unexcelled. It also is often chosen for its high strength at elevated temperatures. Although it can be recommended for a broad range of corrosive conditions its cost often limits its use to only those applications where its exceptional characteristics are required. Hatelloy B-2 is an alloy rich in nickel and molybdenum which has been developed primarily for resistance to reducing acid environments, particularly hydrochloric, sulfuric and phosphoric. Its resistance to these acids in puré form is unsurpassed, but the presence of ferric and other oxidizing ions in quantities as low as 50 ppm can dramatically degrade the resistance of this alloy. Hastelloy C-276 is a nickel chromium-molybdenum alloy having perhaps the broadest general corrosión resistance of all commonly used alloys. It was developed initially for use with wet chlorine, but it also offers excel-

Chapter 8 - Lab Autoclaves & Pilot Plañís

201

lent resistance to strong oxidizers such as cupric and ferric chlorides, and to a variety of chlorine compounds. Nickel 200 is one of the designations of commercially puré nickel. It offers the ultímate in corrosión resistance to host caustic environments, but its applications are severely restricted because of its poor machinability and resultant high fabrication costs. Titanium is an excellent material for use with oxidizing agents, such as nitric acid, aqua regia and other mixed acids. It also offers good resistance to chloride ions. Sulfuric and hydrochloric acids, which have high corrosión rates in their puré form can have their corrosión rates in titanium reduced if small quantities of oxidizing ions, such as cupric and ferric are present to act as corrosión inhibitors. This phenomenon leads to many successful applications where sulfuric acid is used to leach ores and the extracted ions act as corrosión inhibitors. Prospective users must remember that titanium will burn vigorously in the presence of oxygen at elevated temperatures and pressures. While there have been many successful applications in hydrometallurgy where oxygen and sulfuric acid are handled in titanium equipment, the danger of ignition is always present and must be protected against. Commercially puré titanium is available in several grades. Grade 2 is the material most commonly used for industrial equipment since it can be fabricated by welding and is approved by the ASME Code of Unfired Pressure Vessels. Grade 4, which has slightly higher trace levéis of iron and oxygen, has higher strength than Grade 2 but it is not suitable for welding and it is not covered by the ASME Code. Since Parr vessels are not welded, they usually are made of Grade 4 to obtain higher working pressures than can be obtained with Grade 2. Grade 7, containing small amounts of palladium, and Grade 12

202


containing small amounts of nickel and molybdenum, offer enhanced resistance to certain environments and can be used for Parr reactors and pressure vessels if suitable billets can be obtained. Zirconium offers excellent resistance to hydrochloric and sulfuric acids but as with Hastelloy B-2, oxidizing ions such as ferric, cupric and fluorides must be avoided. Zirconium also offers good resistance to phosphoric and nitric acids, and to alkaline solutions as well. Two different grades are available. Grade 702 containing hafnium is the standard commercial grade offering the best resistance to most corrosive agents. Grade 705 containing small amounts of both hafnium and niobium has better strength than Grade 702, allowing higher working pressures when it is used in pressure vessel construction, but the corrosión resistance of Grade 705 is not quite as good as Grade 702. Carbón steel is usually used for laboratory reactors only when it is desired to duplicate construction material used in plant equipment, Because it rusts easily, carbón steel vessels are not carried in stock and must be made to order, often resulting in costs higher than for stainless steel equipment despite the lower material cost for carbón steel.

which assures uniform tightening and a reliable seal by means of two simple hand screws.

Pressure g^ge has stainl«is sfeet tube and socket

Safety ruptura dtsk protects against accidental over-pressure • Compressed gas conneitíon ts made to threaded opemng in valve body

Sttrrer drive turns on balj and needle bearíngs in steet hiib Cones of a Teflon-base píasfic form gas-iíght giand on sfírrer shaft

Gas ínlef valê

Gas reléase *'aive

^ „ ^ L ¡ q u í d sampliiig valv»

Water connecfton to cooling channe! around paclctng gland -— Bomb head is clamped to cyíínder by tightening six cap screws in pair of steeí ring sectíons

Two-piece threaded adapten seal valves and gage fo head, factng ín correct positíon

Heavy steel band holds the clamp ring sectíons in position Thermowelí extends to bottom of cyíínder for femperafure measurenríents with either a thermometer or a thermocoupii

Cyíínder h machined írom soiid hot-roíied bar of T3[6 stainless steel. afío from other corrosíon-resístant metáis tíné alloys Stirrer shaft turns in replaceable Teflon bearíng Gas iníet and ¡tquid samplíng tube

Agitated autoclaves These are versatile laboratory reactors available in 100 mi up to 2 litres from Parr and in large sizes from Autoclave Engineers. They can be quipped with gas inlet tube, cooling coil, agitators, etc. (Figure 8.1). Berghof autoclaves incorpórate an interchangeable, completely pore-free, chemically resistant PTFE lining, covering all interior surfaces. The stirring system does not use a packing gland or magnetic drive; a three-phase induction motor provides reliable stirring, even under heavy load. Reactor operates up to 20 000 kPa and 250'^C. Berghof developed a special conical flange closure

203


Stirrer beartng braclcet is clamped to fhermoweÜ

Two 6-b!ade propeíieri agítate the reactants with a turbulent dowofhrusf. Propellers are adjustabie on stirrer síiaff

Figure 8.1 - Two litres Parr autoclave

TT 204



205

Pressure gauge Stirrer drive system Gas inlet valve Gas reléase valve"

Uquid sampling vaive

Safety rupture disc

Thermocouple

Dlp tube connected to both the gas inlet and liquid sampling valves

Wning shaft Lowerguidebearing

Figure 8.3 - Autoclave head

Figure 8.2 - Cooling coils

J

206


207


\r\ cnJer to carry sway fnciton haal ênerated at IJw packing. the drtve shah is dniled out, srxj by mear» o( £ (otix seal, a cocsteni ttcw oí waier a cinxttatad Bhrough the drive shaíl ín tfis packing se:[ion.

Heawy-duly tíirust bearings insure long-iffe operatlon. Provisión ts (nade lor padung tensión taKe-up. wNch allfjws foF easy maintenance and long t
Oi S6a! hoips pfotocl ihe gauge agalnst corrosivo VJ^KHJTS,

The padtlr>9 hDf standard auiociavfis is micki (rom tcítov-astíestos. Tho coinbinatiofi ol low cooliicwnt ot fricîn and high-corrosion roslstanco próvidos pacKlng wHblowí^raíing tamporalure. long Uto, nigh slrongih. and nnJríimuRi friciionai wear on iho drivo shall. Orive sKait

SoWs clrarging

Figure 8.4 - Bomb closure details The siarxiard ctosure bollad confinad gasKel type

Safety ^ a ü assembly

Electric t>&ating of vapour fackMs

Cooling cotls

Sannpíing A iharmooouptewithtn a th«rmocxHiple

BSow pipe or drain conrMCtion can be pcovided tor emptying Iho autoclavo wtthoul removal oí cover. This is a turblne-type agitator mtn a tvMwí shait used in cotijunciioh wiih rerrMavaW* bafflas.During opecaEion, s low-pfassure atea ts created ai ¡tío aitt>ine Jrt)pellor. Gas is «Dosaquenily drawn dovm Ihrcugh iho hoíiow shan and dJspeised (hroughoui the íiqu«i- Tho gas bubbles are t>rai(«n up by the t>airte^

Figure 8.6 -Autoclave Engineers

Figure 8.5 - Assembled unit

208



I LITER

1 GALLÓN


5 Gallons Figure 8.8 -Autoclave Engineers

J

209

210


211


MOTOR SUfPORT OWG.* i-\/2 H.P MOTX)ft-a. i, GR R 2 2 0 W 4 0 VOLT- 60 CYCLE - 3 PHASC H40 Í?,PM FRAME * L. ALUS MOTOR

ORIVE SHAFT - LOW£R SECTOR 516 S.S. SPACEB- 316 S S , _ RETAiNiniO RiWG

18-1

[16) t-t/4-8 H e x , SOC C0VCT-Sft24O r y P E

CAP SCRS. 3)6 S.S-

SASKET-16-e BO0r-Sfti82

QR-t^Ste

_ AGlTATOR 5HAPT- 316 S.S-

24 STR1P M E A T E R S - I O O O WATTS EACH - WIRED IN 2 BANKS OF 12 K.W, EAW - 220 VOLT s^3 PMASE SERVICE

STAND STEEL

10 Gallons

30 Gallons



212


213


Zipper Clave

msss-xiLjsasmi-

This Autoclave Engineers reactor is manufactured in 2, 1,2 and 4 liters. It is claimed to be the fastest, easiest, opening and closing pressure vessel ever offered. No bolts to torque, no clamps or rings. Instead, just raise the body and push the spring section to cióse the cover. To open, just pulí the spring and lower the body. Cover remains stationary, so there is no need to break cover connections.

ü fíOroB

5W«»C«T

2 «P »OTOB-»00 fiPM ^ ^ • 4 0 «t.T-3 «^Mt to CTCLE a, '.-„'>f O-

Zip . . .

Stationary outer tiousing (no sheaves or boltguard needed)

Insert the cover, zip the Zipper, it's shut.

Un-Zip Í;

Pulí the Zipper, raise the cover, it's open.

1 ^ ^ vm.<«-m»s «TI

500 mi ZípperClave system

Figure 8.13 - 500 mL Zipper Clave system

50 Gallons Figure 8.11 -Autoclave Engineers


214

w


215

Pressure gauge

Vessel handiing devicG (optional) standard on 2 and4litermodels

77ZZ2ZZ^ü^

500 mi model

Figure 8.12 - ZipperClave

•m- MT i r J M ^

II AMERICAN INSTRUMENT CO.

Figure 8.14 - Berghof autoclave

fnzá

j

kcB

Figure 8.15 -American Instrument Co. autoclave

216



217

PRESSURE CAUCE

uuoie DYNA/UAC mXER

MOTOR CONTROLLER

SELF ENERGIZING SEAL

RÍCID ¡SSUUTION

PROCESS TBERUOWELL BVLK mSULATION

REATE R

AGITATOR

SÁMPLE

TUBB

Packing Gland

Pressure Products Industries, Inc.

X

3

2000 psig (138 bar) @ 350° C 300 mi, 500 mi l,2,and4UtersÍ2es. As ihc reactor cover and body are bolicd togeiher, ihe metal sea! ring seats itself. Intemal pressure incieases the seal load and eneigizes the seal ring, assuring a tight. leak-free seal.

Figure 8.16 - Pressure Products Industries Inc. autoclave

Monitoring Nlpple

Coolant Channal

STIRRER ORIVE WITH SELF-SEALING PACKING GLAND

Figure 8.17 - Stirrer drive with self-sealing packing gland

218


Chapter 8 - Lab Autoclaves & Piloi Plants

High torque magnetic drives Using high energy permanent magnets in sealed enclosures with no rotating seáis, these magnetically coupled systems elimínate the troublesome leakage problems which sometimes arise with a packing gland in severe service, permitting long continuous runs at pressures up to 30 000 kPa with little or no attention to the gland and drive. A water cooling sleeve attached to each drive protects the components from excessive temperatures arising from the reactor (Figure 8.X). The stir shaft rotates in PTFE bearings and is rigidly connected to the internal magnet. Both of these components are inside the pressure vessel. The externa! magnet is supported by all bearings and rotates outside the pressure chamber. As a result of the optimally utilized lines of forcé of the magnetic field, the internal magnet rotates in synchronization with the external magnet.

External Driver Magnet Assembly (as seen from above) consists of an outer Steel housing that cotains four equally spaced permanent magnets with attracting poles on their inner faces.

High-torque magnetic drtves

OlJter drivjng magneís

Internal Driver Magnet Assembly is totally encapsulated. It is made up of circular, ceramic-type permanent magnets encasing a square rotor shaft and enclosed in a leak-proof housing.

Inner magnetic rotór in a sealed housing Inner rotor is compíeleíy encfosed wíttiin a non-roíaling housing wiiti fixed seáis.

A water-cooling jacket protects tile magneís and seáis from elevated temperalures.

Stirrer drive shaft

In operatíon, the Driver Magnet Assembly is rotated by a motordriven V-beIt, and the Driven Magnet Assembly rotates in response. This action drives the rotor shaft to which the agitator is connected

Figure 8.19 - MagneDrive Figure 8.18- High-torque magnetic drives

219

w Presstire Hydrometallurgy

220


221

Microwave acid digestión bombs are made of a microwave transparent polymer which has good mechanical strength at temperatures up to 150°C, and which serves also as an excellent heat insulator for the Teñon sample cup. Since heating is developed internally within the cup, temperatures in the outer, high strength body will seldom exceed 50°C. This type of autoclave is characterized by fast digestión times, can opérate at temperatures up to 250°C, and pressures to 8000 kPa.

Acid digestión bombs Parr bombs contain a Teflon cup and are available in'two sizes: 25 mi and 125 mi. They can be heated in an oven. The 125 mi bomb has a safety rupture disc built into the head which protects the bomb and the operator from the hazards of an unexpected explosión from accidental overloading. If the pressure in the bomb should accidentally reach the 3500 psig range, a pair of thin, frangible discs (Inconel + stainless steel) will rupture.

Teflon Screw (pressure indícator) Screw Cap

Berghof bomb is a stainless steel pressure vessel with 10 mi Teflon sample holder, heated from below by a heating coil. Operating temperature up to 180°C.

Compressibie Disc

Closure is effected by resilient spring member (The Zipper) inserted through a circumferential groove in body and cover. Simply reléase the safety sleeve and pulí the Zipper to remove the cover. A quick release/safety lock and cover safety device are provided to ensure that the spring is fully inserted.

Bottom Píate

Figure 8.21 - Microwave acid digestión bomb

Engineering Aspects Vahes Valves used in autoclaves have non-rotating stems to minimize leakage. Because stem tip does not rotate on contact with body seat, the 2-piece Non-Rotating Stems designed by Autoclave Engineers elimínate the galling and scoring of stem and seat usually associated with conventional rotating stem valves (Figure 8.22). Thus, one of the major causes of valve leakage and failure is avoided. Figure 8.20 - Parr acid digestión bomb

M


222


223

ACCESSORIES •XA

i

• DESIGNER HANDLE

íí

PACKiNGGLAND-

SAFETY

L O C K I N G DEVICE

HEAD

STAINLESS STEEL BODY

../: NON-ROTATiNG STEM-ADJUSTABLE PACKINGBELOW THREADS

^ÉÉjMMI i^W^^W

\ ^ r H

l|flt|,|a

¿M^

^1

f|

\iw0^ WK-'

NON-ROTATING STEM VALVE

Figure 8.22 - Non-rotating stem valve

3/16" (4.76 mm) 1/2" ~(1i7mm)—^ DIAM.

FLAT SEAT RUPTURE DISC

" «r~^/^ 1/4" -^(6.35 mm)-^ DIAM.

30° ANGULAR SEAT RUPTURE DISC

Discs can be coated on one or both sides wiíh Teflon. Such coatings and linings increase the mínimum rupture ratings available in a given disc

Accessories Figure 8.23 shows a safety head in which the rupture disc are installed. To prevent the action of corroding vapours, the discs can be coated on one or both sides with Teflon. Such coatings and Uning increase the minimum rupture ratings available in a given disc. Figure shows also a dial thermometer and a connecting tube.

Dial Thermometer

Figure 8.23 -Accessories

224


225


PILOT PLANTS

^^^^^l^^^^^^^^^li^Bi DIGESTIÓN BOMB Handle Screw Cap Pressure Spring Teflon Cap

?

^^xi^zrI I

Stainless Steel Pressure Bomb Teflon Reaction Vessel

Regulation Range:

P

20°C to 220°C

Cooling Chamber

Figure 8.25 -Autoclave assembly unit in operation

HEATING MANTLE Receptacie for Temperature Sensor Heating Coils BERGHOF/AMERICA Figure 8.24 - Digestión bomb and heating mantle

Figure 8.26 - High-pressure hydrometallurgical laboratorv

226


.Chapter 8 - Lab Autoclaves & Pilot Plants

Figure 8.27 - Titanium autoclaves in a hydrometallurgical pilot plant at Colorado School of Mines Research Institute, Golden

PnePAHArtOM

^ TOW

PfiOOUCT

QiSCHARCE

S A M P L E LlWES

*CtO a«llM n£«iajC StttRRV

OXVGEW SUPPLY

Figure 8.28 - CSIRO mini plant

Figure 8.29 - INCO mini plant

227

228


Figure 8.30 - A miniplant with a total operating volume of about 100 litres in four stages, and can opérate at pressures up to 18 bar and temperatures of 180°C. The capacity is 40-60 L/h of feed, or about 7-10 kg/h of sulfide concéntrate. Both oxidative and non-oxidative leaching can be performed [MINTEK, S. África]

Figure 8.31 - Ortest Metallurgical Research, Perth, Australia


Figure 8.32 - Pilot plant units for pressure oxidation in hydrometallurgy [Zeton Canadá]

229

230


231

Index

Ñame index Anthony, M.T. 13 Bayer, KarlJosef 3 Beketoff, Nikolai 2 Collins, Michael 13 Downes, Kenneth 11 Dumas, Jean-Baptiste 2 Flett, Douglas 13 Haber, Fritz 5 Henglein, Friedrich 5

Ipatieff, Vladimir 3 Mackiw, Vladimir 9 Malzac, M. 4 Maslenitsky, ¡van 6, 13 Naboichenko, Stanslav 13 Nemst, Walther 5 Papangelakis, Vladimiros 13 Pawlek, Franz 11 Schaufelberger, Félix 7, 9

Subject index

Acid-resisting brick 22, 32,34Acid digestión bombs 220 Additives 159 Adsorption 55 AkitaZinc 175 Albion process 194 Alliance Copper 183 Alumina from clay 70 Ambatovy 78 American Cyanamid 7 American Instrument Co. 215 Ammonium carbonate 93 Ammonium hydroxide 18 Amur región 130 Anaconda Company 186 Anglo American processes 108 Anodic Slimes 135 Anorthite 68 Anthraquinone 160 Antimodes 87

Aqueous oxidation of sulfides 53 Arbiter process 188 Arsenic disposal 180 Arsenic minerals 140 Arsenides 87, 140 Arsenopyrite 146 Arsine 140 Athabasca Lake 91 Autoclaves 3, 19 Autoclave Engineers 202 B Bacterial leaching 182 BarrickGold 12, 133 Bauxite 3, 60, 68 Bayer process 61 Beaverlodge 93 Berghof autoclave 202, 214 Berghof bomb 220 Bergius process 6 Bingham Canyon 182

252

BIOX 182 Bohmite 60 Boundry layer 49 Bug mili 188 Bulong process 73

Calera Process 9, 142 Canadian Copper Refmers 137 Carbón steel 202 Carlin Trend in Nevada 132 Camotite 90 Carón process 165 Carpenter 20Cb-3 198 Cassiterite 88 Cawse process 73 CESL108 Chem-controlled process 52 Chuquicamata 183 Claude process 6 Clay 68 Clear process 147 Cobalt from pyrrhotite 119 Co-precipitation 54 Cobalt-arsenic ore 145 Coke oven gas 5 Cominco's refinery 112 Copper scrap 166 Crystal growth 54 Cuprex process 192 D Davidite 92 Diaspore 60 Diffusion processes 51


Disproportionation 56 Dominican Republic 126 E Elemental sulfiír 53 Elko, Nevada 133 ElliotLake91

Falconbridge in Timmins 114 Fiberglass autoclaves 149 Fick's law 50 Flash evaporators 38 Fly ash 68 Fort Saskatchewan 98, 163 Forward, Frank 8 Freeport McMoRan 104 G Gallium 67 Galvanox process 191 Geocoat process 184 Gibbsite 60 H Halex process 109 Hatelloy 200 Heat exchangers 41 High torque magnetic drives 21 í Homestake Mining 12 Horizontal autoclave 24 Hudson Bay Mining 114 Hydrargillite see Gibbsite HydroCopper process 190

1

Index

I ilmenite 81 Inconel 600 200 lonic Precipitation 171

Jarosite 174 Joachimsthal 91 K Kaolinite 68 Kazakhmys Corporation 114 Kennecott Copper 188 Key Lake 92 Kittila gold mine 128 Kola Península 69

Laboratory autoclaves 198 Latentes 71 - in Australia 73 - in Madagascar 78 - in Papua New Guinea 76 - in Turkey 80 Leaching in ammoniacal médium 98 Leaching of laterites 22 - of sulñdes 97 Liberating gold from pyrite 124 Lurgi-Mitterberg process 152 M Macraes Goldfield 133 MagneDrive219 Marinduque 165 Materials of construction 198

233

McArther River 92 Mechanically agitated autoclaves 23 Membrane pistón pump 37 Microwave acid bomb 221 Mines Branch 9 Moab 94 Moa latente 9, 71 Molybdenite 117 Molybdenum ion 102 Monel 400 200 Murrin Murrin process 73 N Nepheline 68 New Zealand's gold mines 134 Nickel-cobalt matte 116 Nickel 200 Nickel from pyrrhotite 119 Nickel coins 10 Nickel in Philippines 165 Niobium concéntrales 86 Norilsk process 121 Nucleation 54, 158 O Occlusion 54 Orbite Aluminae 70 Outokumpu Process 138 Outotec 190 Oxyhydrolysis 100

Parr autoclave 203 Parr bombs 220 Peñoles piant 182

234

Ohysical processes 51 Pigment manufacture 81 Pilot Plants 225 Pitchblende 90 Platsol process 119 Pokrovkamine 130 Potassiumjarosite 150 Precipitating sulfides by H^S 26 Precipitation 54 Precipitation by carbón monoxide 169 -byH^S 154, 171 - by reduction 155 -by 80^169 Precipitation of arsenate 178 - ofcopper 165 - of iron oxide 154, 173 Precipitation from non-aqueous médium 167 Pressure leaching plant 19 Pressure Products Industries 216 Pueblo Viejo 126 Pulp density 47 Purification of chalcopyrite 87 Pyrochlore 85 Pyrrhotite - pentlándite 11 R Radium 95 Radon 95 Recovery of leaching 46 Red mud 62 Refractory gold 12 RIS0 National Laboratory 95 Rotating autoclaves 26 Ruhr-Zinkll4, 177


Rutile 82

Safety head 222 San Luis Potosi 193 Scheelite 84 Seeding 3 Selenium 135 Self-sealing packing gland 217 Sepon process 110 Shale 68 Sherritt-Cominco process 152 Sherritt Gordon 8, 98 Shinkolobwe 91 Siderite 66 Silver nitrate 2 Slurry pre-heater 39 Sodium oxalate 66 Solubility of hydrogen 49 Solubility of oxygen 49 Solubility ofsalts 48 Sorelslag 82 Stainless Steel 316 198 Stillwater smelter 116 Synthetic rutile 82

Technical University in Berlin 11 Técnicas Reunidas 192 Telfer project 112 Tellurium 135 Thionates 100 Thiosulfates 100 Thucholite91 Titanium 201

Index

Titanium cladding 35 Transportation of autoclaves 41 Treadwell process 186 Tube autoclave 29 U Uranium ores 23 Uranium oxide 168 Uranium oxides 89

Valves in autoclaves 221 Vanadium in bauxite 68 Vertical autoclave 21 Voisey Bay 122 W Wolframite 84 Wood preservativos 141, 180

Zeton Canadá 229 Zipper Clave system 213 Zirconium 202

235

HIDROMETALURGIA.pdf

Recommend Documents