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Continuous biological leaching of copper from a chalcocite ore and concentrate in a saline environment Weston, Jacqueline May 1995

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CONTINUOUS B I O L O G I C A L L E A C H I N G OF COPPER  F R O M A C H A L C O C I T E ORE A N D C O N C E N T R A T E IN A SALINE E N V I R O N M E N T  by  JACQUELINE M A Y WESTON  B.A.Sc, The University of Toronto, 1990  A THESIS SUBMITTED IN P A R T I A L F U L F I L M E N T OF  THE REQUIREMENTS FOR THE D E G R E E OF  M A S T E R OF APPLIED SCIENCE  in  THE F A C U L T Y OF G R A D U A T E STUDIES  (Department of Metals and Materials Engineering)  We accept this thesis as conforming jifo the required standard  THE UNIVERSITY OF BRITISH C O L U M B I A July 1995 © Jacqueline May Weston, 1995  In presenting this thesis degree  at the  in partial fulfilment  of the  requirements  for an advanced  University of British Columbia, I agree that the Library shall make it  freely available for reference and study. I further agree that permission for extensive copying of this thesis for scholarly purposes department  or  by  his  or  her  representatives.  may be granted It  is  by the  understood  that  head of my copying  or  publication of this thesis for financial gain shall not be allowed without my written permission.  Department of The University of British Columbia Vancouver, Canada  DE-6 (2/88)  ABSTRACT  Biological leaching of a chalcocite ore and its concentrate was studied in both batch and continuous tests in a 3 g/L chloride environment. The feed material was obtained from the Zaldivar mine in Chile. The total copper contents of the ore and concentrate were 1.56% and 17.3% respectively, and the particle sizes were P  80  108 um and P 53 um respectively. The mineralogy consisted mainly of 80  chalcocite (Cu S), pyrite (FeS ) and brochantite ( C u ( S 0 )(OH) ). A mixed bacterial culture 2  2  4  4  6  dominated by Thiobacillus ferrooxidans was obtained from the mine site for use in this study.  Shake flask and batch stirred reactor tests were carried out to determine the amenability of both the ore and the concentrate to biological leaching. The effects of several variables on the rate and extent of copper extraction were investigated. In the shake flask tests and the batch reactor tests, it was determined that the bacterial culture was capable of oxidizing both the ore and the concentrate with average copper extractions of 90% and 97% obtained respectively.  Continuous biological leaching was carried out in a series of three stirred tank reactors (4 L, 2 L and 2 L) for a period of four months. The bacteria were able to adapt to progressively lower system residence times and higher pulp densities. At the lowest residence time achieved of 2% days with a pulp density of 10%, the steady state copper extraction was 82% for the ore and 83% for the concentrate. The copper bioleach rate achieved in the first bioleach stage was 16 mg/L/h for the ore and 250 mg/L/h for the concentrate. This study demonstrated the feasibility of the continuous biological leaching of either a chalcocite ore or a chalcocite concentrate in a 3 g/L chloride environment. The bacteria-rich leach residue from the process would be an excellent source of inoculum for biological heap leaching.  n  Table of Contents Title Page  i  Abstract  ii  Table of Contents  iii  List of Tables  vi  List of Figures  vii  Nomenclature  x  Acknowledgements  xi  Dedication  x  1.0  INTRODUCTION  2.0  L I T E R A T U R E REVIEW 2.1 Microbiological Concepts 2.1.1 Microorganisms 2.1.2 Mixed Bacterial Cultures 2.1.3 Temperature 2.1.4 Nutrients 2.1.5 Inhibitory Substances 2.2 Kinetics of Biological Leaching 2.2.1 Effect of Mineralogy 2.2.2 Effect of Particle Size 2.2.3 Effect of Pulp Density 2.2.4 Effect of Initial Soluble Iron Concentration 2.2.5 Effect of Stirring and Aeration 2.2.6 Adaptation 2.3 Biological Leaching of Chalcocite 2.3.1 Direct Mechanism 2.3.2 Indirect Mechanism 2.3.3 Galvanic Mechanism 2.3.4 Effect of Soluble Iron 2.3.5 pH 2.3.6 Thermodynamics 2.4 Direct versus Indirect Biological Leaching Mechanisms 2.5 Bioleaching of Copper Concentrates 2.6 Continuous Biological Leaching 2.7 Commercial Processes 2.8 Summary  n  1  iii  6 6 6 8 9 9 10 12 13 14 14 16 16 17 19 20 21 22 23 25 26 26 28 29 32 33  3.0  E X P E R I M E N T A L METHODS 3.1 Feed Materials 3.2 Bacterial Culture and Nutrient Media 3.3 Shake Flask Experiments 3.3.1 Apparatus and Procedures 3.3.2 Experimental Design 3.4 Batch Reactor Experiments 3.4.1 Apparatus and Procedures 3.4.2 Experimental Design 3.5 Continuous Reactor Experiments 3.5.1 Apparatus 3.5.2 System Operation 3.5.3 Experimental Design 3.6 pH and Eh Measurements 3.7 Repulp Filtration Method 3.8 Sample Analysis  4.0  RESULTS A N D DISCUSSION 4.1 Shake Flask Experiments 4.1.1 Comparison of Ore and Concentrate Bioleaching 4.1.2 Bacterial Adaptation to the Concentrate 4.1.3 Typical Bioleaching Results 4.1.4 Sterile Control Experiments 4.1.5 Effect of Particle Size 4.1.6 Effect of pH 4.1.7 Effect of Pulp Density 4.1.8 Effect of Initial Soluble Iron Concentration 4.1.9 Effect of Temperature 4.1.10 Summary 4.2 Batch Reactor Experiments 4.2.1 Effect of Pulp Density 4.2.2 Effect of Leach Time on the Extent of Sulphate Formation 4.2.3 Effect of Initial Soluble Iron Concentration 4.2.4 Summary 4.3 Continuous Reactor Experiments 4.3.1 General Results and System Operation 4.3.2 Results at Steady State 4.3.3 Feed Tank Results 4.3.4 Accelerated Short-Circuiting Effects 4.3.5 Ore versus Concentrate Continuous Biological Leaching 4.3.6 Summary 4.4 Acid Consumption  5.0  CONCLUSIONS A N D R E C O M M E N D A T I O N S 5.1 Conclusions 5.2 Recommendations for Further Work at Zaldivar 5.3 General Recommendations iv  :  36 36 37 38 38 39 40 40 42 42 42 45 45 46 46 47 48 48 48 48 50 55 57 58 58 60 62 63 63 63 65 66 67 69 69 73 82 89 90 91 94 98 98 100 101  6.0  REFERENCES  102  APPENDICES A Sample Calculations B Selected Shake Flask Data C Selected Batch Reactor Data D Continuous Reactor Data  109 117 ..123 127  v  List of Tables Table Table Table Table Table Table Table Table Table Table Table Table Table  2.1 2.2 2.3 3.1 3.2 3.3 3.4 3.5 4.1 4.2 4.3 4.4 4.5  - Optimal Pulp Density for the Bioleaching of Copper Sulphide Concentrates - Rest Potentials for Several Sulphide Minerals (Rossi, 1990, Table 2-38) - Summary of Several Continuous Bioleaching Studies - Characteristics of the Zaldivar Ore and Concentrate - Flotation Chemicals Used in the Production of the Zaldivar Concentrate - Shake Flask Experimental Variables - Batch Reactor Experimental Variables - Design of the Continuous Biological Leaching Experiments - Continuous Biological Leaching Steady State Results - Check of Stage 1 Copper Leach Rates - Average and Range of pH and Eh in the Ore and Concentrate Feed Tanks - Summary of Gross Acid Consumption - Percentage of Gross Acid Consumption by Type of Mineral  vi  15 23 31 36 37 40 42 46 74 76 87 94 95  List of Figures Figure 1.1Figure 2.1 Figure 2.2 Figure 2.3 Figure 2.4 Figure 3.1 Figure 4.1 Figure 4.2 Figure 4.3 Figure 4.4 -  Figure 4.5 Figure 4.6 -  Figure 4.7-  Figure 4.8 -  Figure 4.9 -  Figure 4.10 Figure 4.11Figure 4.12 Figure 4.13 -  Figure 4.14 -  Proposed Flowsheet for the Continuous Biological Leaching of a Chalcocite Concentrate at the Zaldivar Mine in Chile A Typical Bacterial Growth Curve (Rossi, 1990, Fig. 2-12) Typical Bacterial Growth Curves Resulting from Adaptation of the Bacterial Culture (Rossi, 1990, Fig. 4-2) The Eh-pH Diagram for the Cu-S-H 0 System at 25 °C The Eh-pH Diagram for the Cu-S-Fe-H 0 System at 25 °C (Ehrlich and Brierley, 1990, Figure 4.2) Schematic Diagram of the Continuous Biological Leaching System Effect of Feed Material and Chloride Concentration on Bacterial Growth in Shake Flasks at a Pulp Density of 10% as Measured by Eh Adaptation of Bacteria to the Zaldivar Concentrate at a Pulp Density of 7.5% in Shake Flask Tests in a 3 g/L Chloride Environment as Measured by Eh . . . . Copper Extraction During the Shake Flask Biological Leaching of the Zaldivar Ore at a Pulp Density of 10% in a 3 g/L Chloride Environment Copper and Iron Concentrations in Solution During the Shake Flask Biological Leaching of the Zaldivar Ore at a Pulp Density of 10% in a 3 g/L Chloride Environment Solution Eh and pH During the Shake Flask Biological Leaching of the Zaldivar Ore at a Pulp Density of 10% in a 3 g/L Chloride Environment Copper Extraction During the Shake Flask Biological Leaching of the Zaldivar Concentrate at a Pulp Density of 10% in a 3 g/L Chloride Environment Copper and Iron Concentrations in Solution During the Shake Flask Biological Leaching of the Zaldivar Concentrate at a Pulp Density of 10% in a 3 g/L Chloride Environment Solution Eh and pH During the Shake Flask Biological Leaching of the Zaldivar Concentrate at a Pulp Density of 10% in a 3 g/L Chloride Environment Copper and Iron Concentrations in Solution During the Shake Flask Biological Leaching of the Zaldivar Concentrate at a Pulp Density of 10% in a 0 g/L Chloride Environment Effect of Ore Particle Size on Sulphate Formation After 5 Days of Shake Flask Biological Leaching in a 3 g/L Chloride Environment Effect of Ore Pulp Density on Biological Leaching in Shake Flask Tests in a 3 g/L Chloride Environment as Measured by Eh Effect of Concentrate Pulp Density on Biological Leaching in Shake Flask Tests in a 3 g/L Chloride Environment as Measured by Eh . Effect of Initial Soluble Iron Concentration on Copper Extraction from the Zaldivar Concentrate at 10% Pulp Density in Shake Flask Tests in a 3 g/L Chloride Environment Effect of Initial Soluble Iron Concentration on the Solution pH During the Shake Flask Biological Leaching of the Zaldivar Ore at 10% Pulp Density in a 3 g/L Chloride Environment 2  4 13 18 27  2  vii  28 43 49 49 51  51 52  52  53  53  54 57 59 59  61  62  Figure 4.15Figure 4.16Figure 4.17Figure 4.18-  Figure 4.19 -  Figure 4.20 Figure 4.21Figure 4.22 Figure 4.23 Figure 4.24 Figure 4.25 -  Figure 4.26 -  Figure 4.27 -  Figure 4.28 -  Figure 4.29 -  Figure 4.30 -  Figure 4.31 -  Figure 4.32 -  Effect of Ore Pulp Density on Copper Extraction During Batch Reactor Tests in a 3 g/L Chloride Environment Effect of Concentrate Pulp Density on Copper Extraction During Batch Reactor Tests in a 3 g/L Chloride Environment Copper and Iron Concentrations in Solution in Duplicate Batch Reactor Tests at a Pulp Density of 5% in a 3 g/L Chloride Environment Effect of Initial Soluble Iron Concentration on Copper Extraction from the Zaldivar Ore at 15% Pulp Density During Batch Reactor Tests in a 3 g/L Chloride Environment Effect of Initial Soluble Iron Concentration on pH During the Batch Reactor Biological Leaching of the Zaldivar Ore at 15% Pulp Density in a 3 g/L Chloride Environment Copper Concentration versus Time for the Ore Continuous Biological Leaching Run in a 3 g/L Chloride Environment Copper Concentration versus Time for the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment Actual Flowrate for the Ore Continuous Biological Leaching Run in a 3 g/L Chloride Environment Actual Flowrate for the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment Copper Leach Rate versus Time in the First Bioleach Stage of the Ore Continuous Biological Leaching Run in a 3 g/L Chloride Environment Copper Leach Rate versus Time in the First Bioleach Stage of the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment Copper Extraction and Sulphate Formation for the Ore Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 8 Days and a Pulp Density of 10% Copper Extraction and Sulphate Formation for the Ore Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 4 Days and a Pulp Density of 10% Copper Extraction and Sulphate Formation for the Ore Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 2% Days and a Pulp Density of 10% Copper Extraction and Sulphate Formation for the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 8 Days and a Pulp Density of 5% Copper Extraction and Sulphate Formation for the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 4 Days and a Pulp Density of 5% Copper Extraction and Sulphate Formation for the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 2% Days and a Pulp Density of 5% Copper Extraction and Sulphate Formation for the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 2% Days and a Pulp Density of 10%  viii  64 64 66  68  68 70 70 72 72 75  75  79  79  80  80  81  81  82  Figure 4.33 - Copper Extraction in the Ore Feed Tank versus the Age of the Feed Tank Slurry Figure 4.34 - Sulphate Formation in the Ore Feed Tank versus the Age of the Feed Tank Slurry Figure 4.35 - Copper Extraction in the Concentrate Feed Tank versus the Age of the Feed Tank Slurry Figure 4.36 - Sulphate Formation in the Concentrate Feed Tank versus the Age of the Feed Tank Slurry Figure 4.37 - Sulphate Formation and the Proportion of Sulphate Formation Attributable to Chalcocite-Sulphide and Pyrite-Sulphide by Stage of the Ore Continuous Bioleaching System for Test 1 Figure 4.38 - Concentrate Continuous Biological Leaching Mass Balance at Steady State with a Residence Time of 2% Days and a 10% Pulp Density  ix  83 83 84 84  86 92  Nomenclature a =  relative activity of component i  D -  impeller diameter (m)  Eh =  redox potential (mV)  TV =  stirrer speed (RPM)  N=  number of bacteria at time t  N=  number of bacteria at time zero  P =  particle size 80% passing (nm)  ;  t  t  Q  80  s.g. = specific gravity SHE = standard hydrogen electrode reference u=  bacterial growth rate (day )  t-  time (days)  v=  impeller tip speed (m/s)  t  1  x  ACKNOWLEDGEMENTS This study was funded by Placer Dome Inc. and the Natural Sciences and Engineering Research Council of Canada. I would like to thank my supervisor, David Dreisinger, for his advice and encouragement. M y sincere thanks are also extended to Ralph Hackl for donating much of his time and expertise to this project. The assistance of Eugene Chong, Anita Lam, Michael Lo and Lawrence Weston in the laboratory is gratefully acknowledged. I would like to thank my parents and my grandparents for encouraging me to pursue a graduate education.  xi  Dedicated to Beaker  xn  1.0  INTRODUCTION  Conventional copper production involves mining of copper ores, followed by milling to extract and concentrate useful minerals, and copper recovery by pyrometallurgical smelting.  A major  disadvantage of smelting is the production of sulphur dioxide (S0 ) gas, which is a contributor to 2  the production of acid rain. Modern smelters are designed to capture most of the S 0 , which is used 2  to produce sulphuric acid. However, increasingly stringent legislated limits on S 0 emissions to the 2  environment have stimulated interest in hydrometallurgical processes in which sulphide is converted to elemental sulphur or sulphate.  These include chloride leaching, ammonia leaching, oxygen  pressure leaching and biological leaching (Biswas and Davenport, 1994; Everett, 1994; Peters, 1992). One advantage of biological leaching is potentially lower capital and operating costs, which are increasingly important as available ore grades decrease. Another advantage is the fact that copper is extracted in sulphate media. More success has been achieved in improving solvent extraction and electrowinning processes with sulphate solutions than with other media such as chloride or ammoniacal solutions (Peters, 1992).  Biological leaching is the bacterially catalysed dissolution of metals from sulphide minerals. The dominant bacteria in biological leaching processes are Thiobacillus ferrooxidans, which live in acidic environments, gaining energy for growth from the oxidation of reduced sulphur and iron compounds. Copper is the metal most commonly extracted by biological leaching. It has been reported that biological leaching accounts for about 5% of the world's primary copper production (Torma et al., 1979) and about 20% of copper production in the United States (Torma and Apel, 1992). The main copper sulphide minerals are chalcopyrite (CuFeS ), chalcocite (Cu S), covellite 2  2  (CuS) and bornite (Cu FeS ). Chalcopyrite is the most abundant copper sulphide mineral but it is 5  4  1  the least reactive to biological leaching. The secondary sulphide minerals, chalcocite, covellite and bornite, are much more reactive to biological leaching than chalcopyrite.  Bioleaching in the laboratory is accomplished in shake flasks, columns, batch reactors and continuous stirred tank reactors.  Commercial biological processes currently used for copper  extraction include in-situ, dump and heap methods. There are currently no commercial continuous stirred tank reactor bioleaching processes for copper extraction, although the process is used commercially for the treatment of refractory gold ores and concentrates. The potential advantages of continuous biological leaching include higher metal recovery, higher copper leach rates and more effective control of the process. The main disadvantage is the cost, which is higher than costs for in-situ, dump and heap bioleaching.  This study considered the continuous biological leaching of a copper sulphide ore and its concentrate from the Zaldivar mine in Chile. Zaldivar is an open pit mine located in the desert region of northern Chile, 196 km to the southeast of Antofagasta and 1400 km north of Santiago. The Compania Minera Zaldivar is a 50% joint venture between Placer Dome Inc. and Outokumpu Copper Resources B . V . The deposit is porphyry copper containing about 50% oxide copper and 50% sulphide copper. There are two main rock zones, one of which is an andesite rock containing either copper oxides, or mixed copper oxide and sulphide minerals. The other is a rhyolite copper porphyry containing mainly supergene sulphide minerals. In the rhyolite zone, which was the source of the feed material for this study, chalcocite (Cu S) makes up about 30 to 60% of the total sulphide 2  content, and about 85 to 90% of the total copper content. The other main sulphide mineral is pyrite (FeS ). The main copper oxide mineral is brochantite (Cu (S0 )(OH) ). 2  4  2  4  6  The Zaldivar copper extraction process includes ore crushing and sizing, heap leaching, and solvent extraction and electrowinning (SX-EW). A fines fraction is separated from the crushed ore, which is about 3% of the total ore mined. When fully commissioned, the annual production will be 125 000 tonnes of cathode Grade A copper. There will be an acid leaching heap for the oxide copper ore and a biological leaching heap for the sulphide copper ore. Dump leaching will contribute a small percentage of copper production. Various options are under consideration for the treatment of the ore crusher fines. One option is to send the fines to a continuous biological leaching system. Another option is to produce a flotation concentrate from the ore crusher fines, and send this concentrate to a continuous biological leaching system. Both of these options were considered in this Master's project.  This study was designed to determine the feasibility of the continuous biological leaching of the Zaldivar ore and concentrate in a saline environment of 3 g/L chloride. A 3 g/L chloride solution was used because the Zaldivar plant water was expected to have this chloride concentration. The Zaldivar ore was proposed to be an excellent application for continuous biological leaching because the main copper sulphide mineral is the secondary sulphide chalcocite, which is known to bioleach more fully and rapidly than other copper sulphide minerals, hi addition, either the ore crusher fines or the fines concentrate would be an ideal feed to a continuous biological leaching circuit. The leachate could be sent to SX-EW for production of copper and the bacteria-rich solids could be used as a source of inoculum for the sulphide heap.  Figure 1.1 gives a schematic diagram that shows how a continuous system for biological leaching of the fines concentrate could fit into the flowsheet of the Zaldivar mine. For continuous biological leaching of the ore crusher fines, the flotation operation shown in Figure 1.1 would not be required. 3  The leaching solution could be sent to SX-EW for copper recovery. The bacteria-rich solid residue from the product slurry could be sent to sulphide heap leaching as a source of bacterial inoculum; addition of these bacteria could increase copper leaching rates in the heap. Some of the copper remaining in the continuous bioleaching system residue could be extracted by heap leaching.  Ore  Crushing and Sizing Coarse  Leach Solution Tailing -*  1 Heap Leaching  Fine riotation Concentrate Residue Used A s Inoculum  i l  Tailing  _ Leach Solution  Continuous Stirred Tank Biological Leaching  Solution to S X - E W Figure 1.1-  Proposed Flowsheet for the Continuous Biological Leaching of a Chalcocite Concentrate at the Zaldivar Mine in Chile  In this study, shake flask, batch reactor and continuous stirred tank reactor biological leaching tests were conducted on the Zaldivar ore and concentrate. The objectives of the shake flask experiments were to verify that the Zaldivar ore and concentrate were amenable to biological leaching, and to  determine the ultimate extent of copper extraction possible. The objectives of the batch reactor experiments were to verify some of the results achieved in the shake flask experiments, and to develop an experimental method for use in the continuous bioleaching experiments. The objective of the continuous reactor experiments was to determine the feasibility of the continuous stirred tank biological leaching of the Zaldivar ore and concentrate in a 3 g/L chloride environment.  5  2.0  LITERATURE REVIEW  Biological leaching is the process in which metals are dissolved from sulphide minerals into aqueous solutions using bacteria as catalysts. The valuable metals can be recovered from the leachate using either conventional technologies such as cementation, or solvent extraction followed by electrowinning. Applications include pretreatment of refractory gold ores and extraction of metals from sulphide minerals. The background and theory of this topic have been covered thoroughly in several books (Barrett et al., 1993; Ehrlich and Brierley, 1990; Rossi, 1990; Karavaiko et al., 1988), and review papers (Ingledew, 1982; Torma and Bosecker, 1982; Murr, 1980). The focus of this review was literature relating to the continuous biological leaching of copper sulphide minerals, particularly chalcocite (Cu S). 2  2.1  Microbiological Concepts  The various bacteria involved in biological leaching and the mechanisms of their growth are described thoroughly in Brock and Madigan (1991) as well as several other sources (Ehrlich and Brierley, 1990; Rossi, 1990; Torma and Bosecker, 1982).  2.1.1  Microorganisms  The most commonly used bacteria in biological leaching processes are the aerobic lithoautotrophs Thiobacillus ferrooxidans. This bacterium was first described after being isolated from acid mine drainage waters in 1947 (Colmer and Hinkle, 1947). They are called aerobic lithoautotrophic bacteria because they require oxygen, inorganic compounds and carbon dioxide for growth. T. ferrooxidans are active in the temperature range of 5 to 45 °C but the optimum temperature range for their growth is 35 °C to 40°C. They require an acidic environment in the range pH 1.4 to pH 6.0, with optimum 6  growth occurring at about pH 2.35 (Temple and Colmer, 1951). These bacteria derive energy from oxidation of sulphur, reduced sulphur and reduced iron compounds. The energy source is oxidized by the bacteria, releasing electrons that are used to drive cellular reactions. T. ferrooxidans are rod shaped with dimensions of 0.5 x 2 microns, and they have a tail-like structure, called the flagellum, that allows motility. Like all bacteria, they have the ability to adapt to changes in their environment. T. ferrooxidans can readily attach to solid surfaces although the mechanism of attachment is not fully understood (Heffelfinger, 1990).  Other bacteria that require similar growth conditions to those described above for T. ferrooxidans are believed to be involved in biological leaching. Examples are Thiobacillus thiooxidans, which oxidize sulphur and reduced sulphur compounds, and Leptospirillum ferrooxidans, which oxidize ferrous iron and some reduced sulphur species.  T. thiooxidans are physically similar to T.  ferrooxidans and they are often found together in nature. However, they can survive in lower pH environments than T. ferrooxidans and they are better at oxidizing elemental sulphur. These bacteria have an optimum pH of 2.0 to 3.5 and grow in the range of pH 0.5 to pH 6.0 (Rossi, 1990, p.45). Leptospirillum ferrooxidans cannot oxidize elemental sulphur and they have a higher tolerance for the ferric ion than T. ferrooxidans. L. ferrooxidans are active in the pH range p H 1.5 to p H 5, with optimal growth at pH 3 (Pivovarova and Golovacheva, 1985). Another mesophilic bacteria that may be involved in biological leaching is the iron oxidizing strain of the filamentous bacteria of the genus Metallogenium. These bacteria exist in the pH range of pH 3.5 to pH 6.8 in the natural environment (Walsh and Mitchell, 1972).  The bacteria described above are known as mesophiles because they are active at temperatures below 40°C. Thermophilic bacteria, which are active in a higher temperature range, may also be present 7  in some biological leaching processes such as in heap or dump leaching. For example, Sulfolobus bacteria oxidize elemental sulphur and ferrous iron, and are active in the temperature range 55 °C to 80°C. However, Sulfolobus bacteria have been found to be susceptible to shear effects during mixing and to date they have only tolerated pulp densities up to 5% (Torres et al., 1995). Moderately thermophilic bacteria also exist, which can oxidize reduced iron and sulphur compounds. These bacteria exhibit optimal growth in the temperature range from about 40°C to about 55 °C. Many moderate thermophiles have not been classified but they are generally thought to belong to the genus Sulfobacillus. A moderately thermophilic bacterial culture is being used at the Youanmi mine in Australia for the biooxidation of a refractory gold concentrate in a continuous stirred tank reactor system (Rhodes et al., 1995).  2.1.2  Mixed Bacterial Cultures  Mixed cultures of bacteria can achieve higher rates of biological leaching than pure cultures of T. ferrooxidans (Norris and Kelly, 1978). One reason for this is thought to be the presence of heterotrophic bacteria in mixed cultures, which metabolize organic compounds that would otherwise inhibit the growth of the lithoautotrophic bacteria (Groudev, 1985; Harrison et al., 1980). Examples of heterotrophic bacteria are the Acidiphilium species and Thiobacillus acidophilus. Heffelfinger (1990) studied the adherence oi Acidiphilium cryptum to chalcopyrite and found that there was no competition between these bacteria and T. ferrooxidans for sites on the mineral surface. Another reason for the improved leaching by mixed bacterial cultures may be a symbiotic relationship that exists between different lithoautotrophic bacteria.  For example, it was found that none of  Leptospirillum ferrooxidans, T. organoparus, T. thiooxidans or T. acidophilus could leach pyrite on their own. However, mixtures of L. ferrooxidans with one of the other three organisms leached pyrite at the same or higher rates than a pure culture of T. ferrooxidans (Norris and Kelly, 1978). 8  2.1.3  Temperature  Malouf and Prater (1961) found that temperature had a significant effect on bacterial oxidation of ferrous ion by T. ferrooxidans, with maximum activity occurring at 35 °C. They also reported that the bacteria were killed at temperatures over 50°C. Sakaguchi et al. (1976a, 1976b) found that the optimum temperature for bioleaching of chalcocite (Cu S) and covellite (CuS) was 35 °C. 2  2.1.4  Nutrients  Nutrients for bioleaching are provided in the laboratory in defined media such as the 9K solution, which contains the nutrients ammonium, phosphate, magnesium, potassium, calcium and ferrous iron (Silverman and Lundgren, 1959). Such nutrients are not typically added to industrial scale bioleaching processes because they are available from gangue minerals. If necessary, ammonium and phosphate can be added in the form of fertilizers. In addition to these nutrients, oxygen and carbon dioxide must be supplied for bacterial growth.  The oxygen supply should be higher than that required for growth of T. ferrooxidans alone, due to the requirements from other organisms that do not contribute directly to the bioleaching process. Dissolved oxygen was found to be rate limiting below 0.3 mg/L during bioleaching, and below about 0.2 mg/L dissolved oxygen the bacteria began to die (Liu et al., 1988). However, during the bioleaching of a gold concentrate, Pinches et al. (1988) found that dissolved oxygen was not rate limiting at levels as low as 0.1 to 0.5 mg/L. In order to ensure that dissolved oxygen is not rate limiting, Gormely and Branion (1989) recommended maintaining dissolved oxygen at 1 to 2 mg/L.  In 1971, a stirred tank reactor bioleaching system was patented in which the optimal enrichment of C 0 in air was reported to be 1% (Duncan and McGoran, 1971). Another study showed that 0.2% 2  9  C 0 in air was adequate but 1% C 0 in air gave enough excess that dissolved C 0 2  2  2  was not rate  limiting (Torma et al., 1972). Nagpal studied the effect of C 0 enrichment of the air supply to a 2  continuous bioreactor leaching a refractory pyrite arsenopyrite gold concentrate. He found that the optimal cell growth rate was achieved with 0.5% to 1 % C 0 in air and that cell growth rate decreased 2  sharply with no C 0 enrichment (Nagpal, 1993). 2  2.1.5  Inhibitory Substances  A substance is inhibitory to bacteria i f its presence causes the lag time to increase and the bacterial growth rate to decrease. As will be discussed in Section 2.2.6, bacteria can be adapted to inhibitory substances such as copper, organic chemicals and chloride.  Malouf and Prater (1961) found that although T. ferrooxidans could be adapted gradually to increasing concentrations of metal ions, they were sensitive to sudden increases in metal ion concentration.  Various researchers have shown that it is possible to adapt bacteria to copper  concentrations of up to 55 g/L in biological leaching systems (Bruynesteyn and Duncan, 1971; Sakaguchi et al., 1976b; McElroy and Bruynesteyn, 1978; Lawrence et al., 1984; Groudev, 1985). Dave et al. (1979) were successful in adapting T. ferrooxidans to 64 g/L copper.  Some organic chemicals can be inhibitory or toxic to lithotrophic bacteria such as T. ferrooxidans. Potential sources of organic chemicals in the bioleaching process are flotation chemicals, solvent extraction reagents and waste products of bacterial metabolic reactions. Certain flotation reagents were of particular interest in this study, including Dowfroth 250 and potassium amyl xanthate. Work by Tuovinen (1978) on several flotation reagents suggested that the inhibition effect was actually caused by the products of decomposition of the flotation reagents in the low pH environment. In 10  particular he found that Dowfroth 250 had a low degree of inhibition, and potassium amyl xanthate had a relatively high inhibitory effect on ferrous ion oxidation (Tuovinen, 1978). The toxicity of weak organic acids to T. ferrooxidans maybe due to their diffusion into and accumulation in the cell, which is caused by the pH gradient between the cytoplasm and the cell's environment (Ingledew, 1982). Alexander et al. (1987) found that low molecular weight organic acids accumulated in the cytoplasm of T. ferrooxidans and attributed the inhibitory effect to either the high concentration in the cell cytoplasm or their acidification of the cell cytoplasm. Gentina et al. (1987) studied the effect of various flotation reagents and solvent extractants on the biological leaching of copper sulphide ore using a pure strain of T. ferrooxidans. They did not find any inhibitory effect from the solvent extractants.  However, they found that the flotation reagents, which included Dowfroth 250,  significantly inhibited bacterial growth. Microbial growth rate was more inhibited than the rate of oxidation of ferrous to ferric ion, suggesting that in the presence of flotation reagents the bacteria required more energy for maintenance of cellular functions and had no extra energy for synthesis of new cells. Hackl et al. (1989) tested several flotation reagents and concluded that since many flotation reagents decompose with time in acid environments, the best way to determine effects of flotation concentrates on a commercial scale is to carry out continuous tests using fresh concentrate.  In general, Thiobacilli bacteria have been found to be sensitive to anions such as the chloride ion. Cameron et al. (1984) studied the effects of salinity on biological iron oxidation and found that the bacteria could be partially adapted to 9 g/L chloride but they could not be adapted to 18.2 g/L chloride (3% w/v NaCl), which is the typical chloride concentration in sea water. They were unable to isolate iron-oxidizing bacteria from sea water and could not find reports of a marine bacterial iron oxidizer in the literature. The inhibitory effects of inorganic anions such as chloride are dependent on the p H of the medium and are likely due to acidification of the bacterial cell cytoplasm. For 11  example, the chloride concentration required for inhibition of iron oxidation was 15 times lower at pH 0.94 than at pH 3.0 (Alexander et al., 1987). Leong et al. (1993) were able to adapt a mixed culture of Thiobacilli growing on a chalcocite ore to 5 g/L chloride in shake flask and column bioleaching tests.  2.2  Kinetics of Biological Leaching  The bacteria involved in biological leaching grow by using the electrons released during the oxidation of reduced iron and sulphur compounds to drive cellular reactions. In particular the electrons are used at the inner cell wall surface to reduce oxygen for the production of water. This allows the cell to maintain a neutral internal pH despite its acidic environment. Bacterial growth is typically first order according to Equation 1, where N is the number of bacteria at a particular time, t  N is the number of bacteria at time zero, u is the growth rate (day ) and t is time (days): 1  0  N  t  = N e« 0  (1)  Torma and Apel (1992) studied the kinetics of the biological leaching of a chalcopyrite concentrate and found that the reaction was first order. Figure 2.1 gives a typical bacterial growth curve showing the various phases of bacterial growth in batch mode. Continuous biological reactors are generally operated in the exponential phase of bacterial growth.  The growth rate of a bacterial culture is dependent on environmental conditions, such as pH, temperature, nutrient concentrations and the presence of inhibitory substances. These factors were discussed in Section 2.1. Other factors can also affect the bacterial growth rate in a biological leaching process, including the following: 12  •  nature of the substrate (eg. mineralogy, metal content),  •  mineral surface area (particle size and pulp density),  •  initial soluble iron concentration,  •  stirring and aeration,  •  adaptation.  2.2.1  Effect of Mineralogy  Secondary copper sulphide minerals such as chalcocite and covellite are more amenable to biological leaching than primary minerals such as chalcopyrite. In a continuous biological leaching study, copper extractions of over 90% were achieved with copper concentrates containing mainly secondary 13  sulphide minerals, while copper extractions of about 50% were achieved with those consisting of mainly chalcopyrite (Groudev, 1985). Low copper extractions achieved during the bioleaching of chalcopyrite are often attributed to the precipitation of jarosite on the surface of the mineral, which prevents further leaching (Groudev, 1985). A process for the silver catalysed bioleaching of chalcopyrite, which improves copper extraction and eliminates costly regrind steps, has been developed (Bruynesteyn et al., 1986; Lawrence et al., 1984; Short and Parkinson, 1983).  2.2.2  Effect of Particle Size  Both the rate and extent of metal extraction have been reported to be functions of mineral particle size (Bruynesteyn and Duncan, 1971). Particle size is particularly important during the bioleaching of chalcopyrite. As discussed in section 2.2.1, it is often necessary to regrind chalcopyrite in order to achieve reasonable metal extraction. It has been shown that decreasing the particle size increases the rate of leaching (Torma et al., 1972; Sanmugasunderam, 1981; Blancarte-Zurita, 1988).  Mineral surface area, which is a function of both mineral particle size and pulp density, is also an important parameter in biological leaching. Pinches et al. (1976) found that during the batch tank bioleaching of a chalcopyrite concentrate, the rate of bioleaching was dependent on the available mineral surface area. They suggested that to increase the leach rate one could either, decrease the mineral particle size or increase the pulp density. The importance of mineral surface area was also stressed by Gormely et al. (1975).  2.2.3  Effect of Pulp Density  Pulp density is the mass of solids per unit volume of slurry. There is an optimal pulp density in biological leaching systems, which allows the highest possible throughput without a reduction in 14  leach rate. The leach rate has been found to decrease at high pulp densities due to limited supply of carbon dioxide to the bacteria (Pinches et al., 1976). A build-up of inhibitory factors may also limit bacterial growth at high pulp densities. Since the optimal pulp density depends on the characteristics of the particular feed material, it is determined experimentally. The optimal pulp density determined for the biological leaching of various copper sulphide concentrates and the corresponding copper leach rates are given in Table 2.1.  Table 2.1 - Optimal Pulp Density for the Bioleaching of Copper Sulphide Concentrates Mineralogy  Optimal Pulp Density  Reference  (%)  Copper Leach Rate (mg/L/h)  2  shake flask  22  215  Sakaguchi et al., 1976b  CuFeS  2  shake flask  4 to 5  5 to 17  Babij andMadjwick, 1983  CuFeS  2  shake flask  5  10 to 36  Briceno and Rossi, 1987  shake flask  5  42  Briceno and Rossi, 1987  batch reactor (3.5L)  -5  25  Pinches et al., 1976  continuous reactor (1-stage, 30L)  40  >400  McElroy and Bruynesteyn, 1978  continuous reactor (5-stage, 5 L each)  25  250 to 410  Groudev, 1985  continuous reactor (5-stage, 5 L each)  20  430 to 470  Groudev, 1985  CuFeS  Cu S + CuFeS 2  CuFeS  2  CuFeS  2  CuFeS  2  Cu S + CuFeS 2  Type of Experiment  2  2  15  2.2.4  Effect of Initial Soluble Iron Concentration  Since ferrous iron is a source of energy for T. ferrooxidans, soluble iron concentration is an important parameter in bioleaching systems. Copper bioleach rates from museum grade covellite in shake flask experiments at 6% pulp density, were found to increase in the presence of 9 g/L soluble iron relative to copper bioleach rates achieved with no soluble iron present (Groudev, 1980). The increases in copper bioleach rates ranged from 5% to 15% for various strains of T. ferrooxidans. Briceno and Rossi (1987) also studied the effect of varying the initial soluble iron concentration in shake flask bioleaching experiments. They found no significant effect on the lag period or the rates of copper leaching when the initial ferrous iron concentration varied from zero to 2.8 g/L.  2.2.5  Effect of Stirring and Aeration  Higher rates of biological leaching are achieved by stirred batch reactor bioleaching than by shake flask bioleaching (Groudev, 1985). Improved stirring and aeration in batch reactors results in higher rates of oxygen transfer to the mineral surface, higher carbon dioxide transfer to the bacteria, and better transport of other nutrients to the bacteria.  Increasing the rate of stirring increases particle suspension and aeration, but it also increases the shear forces acting on the cells. Bacteria are typically not considered sensitive to shear forces in stirred reactors due to their size, which is less than the microscale of turbulence in a typical bioreactor, and the strength of their cell walls (Thomas, 1993). There are two main types of stirrer impellers: radial flow and axial flow. Radial flow impellers, such as the Rushton turbine, have high shear rates and are better at air dispersion than axial flow impellers. Axial flow impellers, such as the pitched blade type, have lower shear rates and are better at particle suspension. Hackl et al. (1989) tested the shear effect of both radial and axial impellers in a bench scale continuous 16  bioleaching experiment and found that bacteria were sensitive to shear effects produced by radial flow impellers, but there was no observable effect when axial flow impellers were used.  Impeller tip speed (v,) is often used as a measure of shear rates in stirred tanks because it has been found not to vary much during scale-up of bioreactors. It is given in Equation 2 in units of m/s where N is the stirrer speed in R P M and D is the impeller diameter in metres: {  v,  = izND. (—) 60 1  V  n\  ( Z )  A typical value for v in industrial fermenters is 5.5 m/s (Einsele, 1978). It is important to note, t  however, that Equation 2 does not consider the type of impeller.  2.2.6  Adaptation  The bacterial cell can be thought of as a "super chemical reactor" because, unlike regular chemical reactors, it can improve its efficiency by genetically adapting to its environment. Bacteria have part of their genetic material on structures called plasmids, which can be transferred quickly and easily between cells. Plasmids typically contain genes for auxiliary cellular functions such as optimum temperature for growth or resistance to metals (Martin et al., 1981). Therefore, bacteria can be adapted to increased concentrations of inhibitory substances such as metal ions, organic chemicals or chloride. This adaptation process can occur naturally or it can be accomplished in the laboratory. For example, cells growing in the laboratory in media containing a certain concentration of the inhibitory substance can be transferred to new media containing a slightly higher concentration of the substance. The bacteria are then cultured at the new concentration until they are adapted. The process can then be repeated until the desired concentration is achieved. The lag phase decreases 17  and the growth rate increases as bacteria become adapted to a substrate (Elzeky and Attia, 1989). Typical bacterial growth curves resulting from an adaptation experiment are shown in Figure 2.2. Different strains of T. ferrooxidans from various environmental conditions have been found to contain different genes on their plasmids, and have varying physical and chemical characteristics such as optimum temperature for growth and resistance to heavy metals (Leduc and Ferroni, 1993). Therefore, the most efficient bacterial culture for bioleaching a particular ore is most likely to be one that has already been adapted to the environmental conditions at the mine.  time  Figure 2.2 -  Typical Bacterial Growth Curves Resulting from Adaptation of the Bacterial Culture (Rossi, 1990, Fig. 4-2)  Genetic engineering techniques may also be applied to develop bacteria that are adapted to the desired conditions. However, a genetically engineered strain of T. ferrooxidans has not yet been developed. T. ferrooxidans genes have been transferred to E. coli bacteria, but a means of returning modified genes to T. ferrooxidans has not been determined (Ehrlich and Brierley, 1990, p.29). 18  2.3  Biological Leaching of Chalcocite  Sizable ore bodies of chalcocite have been reported to occur in Chile, Mexico, Namibia, Peru, Russia and the U S A (Mottana et al., 1978). Over the past thirty years, various researchers have studied the biological leaching of chalcocite. Duncan and Trussell reported that over 90% of the copper in museum-grade chalcocite could be extracted by biological leaching for 30 days (Duncan and Trussell, 1964). Biological leaching of synthetic chalcocite using T. ferrooxidans was studied by Sakaguchi et al. (1976a), who found that the extent of leaching of chalcocite was 90% to 100%. The feasibility of in-situ leaching of a chalcocitic ore was studied (Johnson et al., 1988). Column biological leaching of chalcocite bearing ores has been reported (Leong et al., 1993), and a mineralogical study of the bioleach residues was carried out (Melluish et al., 1993).  The oxidation of copper sulphide minerals at ambient temperature and pressure occurs chemically but the kinetics are slow. Bacteria are known to catalyse these reactions by three mechanisms: direct, indirect and galvanic oxidation. Chalcocite oxidation, whether chemical or biological, occurs in two major steps (Nielsen and Beck, 1972). The first step is the conversion of chalcocite (Cu S) to 2  covellite (CuS), and the second step, which is kinetically slower, is the oxidation of covellite (Beck, 1977). It is unclear what exactly is oxidized during the conversion of chalcocite to covellite. In studying chalcocite biooxidation, Nielsen and Beck (1972) concluded that the apparent source of energy for carbon dioxide fixation was provided by the oxidation of cuprous to cupric ions. However, there is no direct evidence that T. ferrooxidans oxidize cuprous copper. Most sulphide minerals are semiconductors with covalent bonds, in which electrons are shared between atoms. Therefore, it is possible that the chalcocite structure as a whole is oxidized to covellite (Beck, 1977).  19  2.3.1  Direct Mechanism  Direct oxidation occurs when the bacteria attach themselves to the mineral surface and oxidize the mineral using oxygen as the oxidant. Biological oxidation of chalcocite by the direct mechanism is shown by the following two reactions:  Cu S + 0.5O + H S0 2  2  2  -  4  CuS + 20  CuS0  -  2  + CuS + H 0  4  2  CuS0  (3)  (4)  4  These reactions occur spontaneously, but the kinetics are slow at ambient temperature. It was found that bacterial action increases the rate of chalcocite oxidation, as shown in Equation 3, by about forty times (Beck, 1977). Chalcocite oxidation is kinetically faster than covellite oxidation (Beck, 1977).  Iron sulphide minerals such as pyrite (FeS ) are often found with copper sulphide minerals. 2  Biological oxidation of pyrite (FeS ) by the direct mechanism is shown in Equation 5: 2  2FeS + 7.50 + H 0 2  2  2  -> Fe (S0 ) 2  4  3  + H S0 2  4  (5)  This reaction is important to the biological oxidation of copper sulphide minerals for two reasons. Firstly, since pyrite is a major component of many sulphide ore bodies, including copper sulphide ore bodies, pyrite oxidation is the prime sulphuric acid generator. Secondly, pyrite oxidation is a possible source of soluble iron, which is involved in the indirect mechanism as will be described in Section 2.3.2. It is important to note, however, that copper sulphide minerals leach preferentially to pyrite because the latter is more noble (Dutrizac and MacDonald, 1973).  20  2.3.2  Indirect Mechanism  In the indirect mechanism, the bacteria do not attach to the mineral surface but are suspended in the leach solution, where they catalyse the oxidation of ferrous to ferric ion. This results in the regeneration of the ferric ion, which acts as the oxidant for the leaching of chalcocite:  2FeS0 + 0.5O + H S0 4  2  2  -  4  Fe (S0 \ 2  + H0  4  (6)  2  Iron is usually present in solution from the bacterially catalysed oxidation of pyrite, as shown in Equation 5 above. Chalcocite oxidation by the indirect mechanism occurs in two major steps, just as it does by the direct mechanism. The reactions for the oxidation of chalcocite to covellite by the indirect mechanism, and the oxidation of covellite by the indirect mechanism are:  Cu S + Fe (S0 ) 2  2  4  3  CuS + Fe (SO \ 2  A  -»  CuS + CuS0 + 2FeS0  (7)  -  CuS0  (8)  4  4  + S° + 2FeS0  4  4  The oxidation of chalcocite is kinetically faster than the oxidation of covellite (Beck, 1977).  As shown in Equation 8, covellite is oxidized to copper sulphate and elemental sulphur by the indirect mechanism. Bacteria then catalyse the oxidation of elemental sulphur to sulphate by the direct mechanism:  S° + 1.50  2  + H0 2  -  H S0 2  4  (9)  Elemental sulphur is thermodynamically unstable under the conditions of biological leaching. 21  However, at ambient temperature and pressure, elemental sulphur occurs as S , which is a very stable 8  eight-membered sulphur atom ring (Vaughan and Craig, 1978). Therefore, the kinetics of elemental sulphur oxidation are slow at ambient temperature and pressure.  2.3.3  Galvanic Mechanism  Galvanic oxidation occurs when two sulphide minerals with different rest potentials are in contact in solution and an electrical current results between the two minerals. The less noble sulphide acts as an anode and is oxidized. The more noble sulphide acts as a cathode and receives the electrons, which are used for reduction of oxygen gas to water on the mineral surface.  The galvanic  mechanism results in preferential leaching of the less noble sulphide mineral because the rate of leaching of the less noble mineral is accelerated, while the rate of leaching of the more noble mineral is retarded. Galvanic oxidation occurs chemically but it is enhanced by the presence of bacteria. Bacteria act as catalysts in galvanic oxidation by oxidizing the elemental sulphur that is produced in the galvanic oxidation process. Galvanic oxidation can occur during biological leaching in a stirred tank reactor because contacts between sulphide minerals are made and broken continuously. It is expected that increasing the pulp density in a stirred tank reactor would increase the occurrence of galvanic oxidation because of increased mineral contact.  For example, the leaching rate of chalcopyrite in acid ferric sulphate solutions is accelerated in the presence of a more noble sulphide such as pyrite (Karavaiko et al., 1988; Dutrizac and MacDonald, 1973). Dave et al. (1979) found that sphalerite (ZnS) leached preferentially to covellite (CuS), and in fact there was no copper dissolution in the presence of ZnS. During the continuous bioleaching of a chalcopyrite and pyrite bearing concentrate, pyrite oxidation did not occur despite 80 to 90% chalcopyrite oxidation (Lawrence et a l , 1984). Almendras et al. (1987) studied the biological 22  leaching of a copper sulphide concentrate consisting of mainly chalcopyrite and pyrite with some chalcocite and covellite also present. They found that chalcocite leached first, followed by covellite, chalcopyrite and then pyrite. During the batch reactor bioleaching of a gold-bearing arsenopyritepyrite concentrate, it was observed that arsenopyrite began leaching before pyrite and was almost fully leached before pyrite leaching commenced (Miller and Hansford, 1992).  Rest potential is a basic property of semiconducting minerals such as sulphides. It is the observed potential on an open circuit and it is a measure of the nobility of the mineral. Relative rest potentials can be used to predict the leaching behaviour of a mixture of sulphide minerals (Sato, 1966; Jyothi et al., 1988). The rest potentials of several sulphide minerals are given in Table 2.2.  Table 2.2 - Rest Potentials for Several Sulphide Minerals (Rossi, 1990, Table 2-38) Mineral  Rest Potential (mV, SHE)  pyrite (FeS )  540  2  chalcopyrite (CuFeS )  434  covellite (CuS)  434  chalcocite (Cu S)  416  sphalerite (ZnS)  181  2  2  2.3.4  Effect of Soluble Iron  The concentration of iron in solution is very important in biological leaching because of the role of the ferric ion in the indirect mechanism. At elevated pH or when the concentration of ferric iron in solution is above a certain concentration, precipitation of jarosite or ferric hydroxide can occur. The 23  precipitation of hydronium jarosite occurs as follows:  2>Fe {S0 \ + \4H 0 2  4  2H OFe (S0 ) (OH)  2  3  3  4 2  6  + 5H S0 2  4  (10)  In addition, sodium, potassium and ammonium jarosites can precipitate out of solution. Ferric hydroxide precipitates can also form, as shown in Equation 11:  Fe (S0 \ 2  4  + 6H 0 2  2Fe(OH)  3  + 3H S0 2  4  (11)  Sakaguchi et al. (1976a) found that jarosite precipitation occurred above soluble iron concentrations of 0.22 g/L to 0.56 g/L for chalcocite bioleaching at pH 1.7, and above 0.22 g/L to 1.12 g/L for covellite bioleaching at pH 2.3. During the biological oxidation of ferrous sulphate, it was found that ferric hydroxides precipitated out first, followed by jarosites, but by the end of the process the precipitates were found to be almost all jarosites (Toro et al., 1988). A study on the characterization of leach residues from column leaching of a chalcocitic ore found that dominant precipitates were jarosites and basic iron oxides (Melluish et al., 1993). Wiertz et al. (1994) found that ferric ion precipitation was enhanced by the presence of bacteria and also by the presence of solids. They found that jarosite was formed preferentially over ferric hydroxide in the presence of solids. Ferric iron may have a buffering effect in bioleach systems by maintaining the p H at the low values required for the growth of T. ferrooxidans (Duncan and Trussell, 1964). Wiertz et al. (1994) found that ferric ion in the leaching solution helped maintain solution pH at low values. A t elevated pH 24  values, the equilibrium of Equations 10 and 11 would shift to the right, generating acid and causing the pH to decrease. The solubility of the ferric ion is extremely low at pH values greater than 2.5. Jarosites and ferric phosphate precipitates were found to occur at pH values as low as 1.35 to 1.5 in a packed-bed bioreactor meant for iron oxidation (Grishin et al., 1988).  The disadvantages of forming jarosites are formation of insoluble coatings on mineral surfaces that reduce leaching rates, plugging of heap and dump leach beds, and removal of nutrients such as sodium, potassium and ammonium from solution. Leach residues containing jarosites are a disposal problem because they are bulkier than the slag produced by smelters, and they are less stable because they can drain acid solutions containing heavy metals (Peters, 1992).  Since T. ferrooxidans oxidize ferrous iron to ferric iron, at high levels of biological activity most of the soluble iron has been oxidized to the ferric state. Thus a high potential indicates a high degree of bacterial activity. Therefore, redox potential (Eh) is often used as an indirect measure of the level of bacterial activity in biological leaching processes. The redox potential of an active bacterial culture is about 700 mV (SHE) to 900 mV (SHE), which is the potential range of the ferric/ferrous couple depending on the ferric to ferrous ratio. The Nernst equation giving the redox potential for the ferric/ferrous couple at 25 °C is shown below:  Eh  = 0.771  + 0.0591 log  aFe +3 aFe +2  2.3.5  (12)  pH  The p H will increase initially during the biological oxidation of a chalcocitic ore or concentrate 25  (Nielsen and Beck, 1972; Silver and Torma, 1974; Beck, 1977; Groudev, 1985). This is due to the conversion of chalcocite to covellite by the direct mechanism and the oxidation of ferrous sulphate, which are both acid consuming reactions. Later, the pH will decrease due to the oxidation of pyrite, the oxidation of elemental sulphur and the formation of ferric ion precipitates, which are all acid generating reactions (Groudev, 1985).  2.3.6  Thermodynamics  The Eh-pH diagram for the copper-sulphur-water system is given in Figure 2.3. This diagram was generated by Outokumpu's HSC Chemistry computer program. It shows that chalcocite (Cu S) and 2  covellite (CuS) are thermodynamically stable in the pH range 1 to 2 when the Eh is less than about 400 mV (Standard Hydrogen Electrode reference, SHE). However, they are thermodynamically unstable in most active bioleaching solutions, which have a p H of about 1 to 2 and an Eh of about 700 to 800 mV (SHE). The Eh-pH diagram for the copper-sulphur-iron-water system is given in Figure 2.4. The dashed box in this diagram gives the stability and activity limits for T. ferrooxidans. The filled and open circles in Figure 2.4 indicate measured Eh values during chalcopyrite leaching in the presence and absence of bacteria respectively (Ehrlich and Brierley, 1990, p. 84). The bacteria grow in the region where the products of sulphide oxidation are thermodynamically stable and sulphide minerals are thermodynamically unstable.  2.4  Direct versus Indirect Biological Leaching Mechanisms  The relative importance of the direct versus the indirect mechanism of biological leaching is not clear. Support for the direct mechanism includes the fact that T. ferrooxidans can oxidize chalcocite even in the absence of iron (Nielsen and Beck, 1972; Imai, 1978; Dave et al., 1979). The leaching of a chalcopyrite concentrate was found to be significantly enhanced in the presence of bacteria as 26  Eh (Volts) 1.0  Cu-S-H20 1  1  1  1  - System at25.00 C i i i !"" i - . J  i  i  r """ i  i  0.8 0.6  _  0.4  -  0.2  -  '  Cu(+2a)  -  ,  CuO  -  CuS  0.0  -0.2 -0.4  ^~~r~————^Tr^^T^^  -0.6 -0.8 -1.0  1  -2  i 0  File:  i  i  i  i  i  i  i  i  i  i  10  8  2 4 C:\HSC\CUS2500.IEP  Figure 2.3-  -  Cu2S  Cu  _  CM  i  12  14 pH  The Eh-pH Diagram for the Cu-S-H 0 System at 2 5 ° C 2  compared to ferric sulphate leaching, suggesting that the direct mechanism plays an important role in biological leaching (Almendras et al., 1987). Vargas et al. (1990) evaluated the numbers of attached and free T. ferrooxidans during the biological leaching of pure chalcopyrite and found that a significant fraction of the total bacteria in the system were attached to the mineral surface. However, Espejo and Ruiz (1987) studied the ratio of attached bacteria to bacteria in solution and found that less than 10% of the total bacterial activity, as measured by ferric iron generation, was associated with attached bacteria. Therefore, although it is clear that both direct and indirect mechanisms play a role in biological leaching, the relative importance of each is not known.  27  TO  Figure 2.4 -  2.5  The Eh-pH Diagram for the Cu-S-Fe-H 0 System at 25 °C (Ehrlich and Brierley, 1990, Figure 4.2) 2  Bioleaching of Copper Concentrates  Much of the interest in copper sulphide biological leaching has been focussed on copper extraction from low-grade ores (Torma and Bosecker, 1982; Murr, 1980). However, there is also much interest in biological leaching of copper concentrates. The feasibility of biological leaching chalcopyrite concentrates has been demonstrated (McElroy and Bruynesteyn, 1978; Torma et al., 1979; Babij and Madgwick, 1983; Groudev, 1985; Agate and Khinvasara, 1986; Torma and Apel, 1992; Corral et al., 1993). Bioleaching of chalcocite bearing copper concentrates has also been demonstrated (Groudev, 1985). Biological leaching of copper concentrates can be more complex than biological 28  leaching of copper ores. With concentrates, copper contents are typically higher and the presence of flotation chemicals can inhibit bacterial growth. The copper content of concentrates varies from about 15-40 wt% total copper and bacteria must be acclimatized to the high copper levels. Since each concentrate has a unique mineral composition, it is necessary to test individual concentrates to determine their amenability to biological leaching.  2.6  Continuous Biological Leaching  Continuous biological leaching is typically carried out in a series of stirred tanks operated at steady state. The bacterial cultures are maintained in the exponential growth phase in order to achieve the highest possible leaching rates. Therefore, the lag time associated with batch bioleaching is eliminated. Continuous bioleaching rates are typically higher than batch leaching rates achieved with the same feed materials (Groudev, 1985; Corral et al., 1993). A n important design consideration is bacterial washout, which can occur if the system residence time is less than the doubling time of the bacteria. To reduce the chance of washout, the residence time of the first stage should be greater than the residence time of the subsequent tanks (Gormely and Branion, 1989).  This can be  accomplished by increasing the size of the first tank or by dividing the first stage into two or more equal sized tanks operating in parallel.  hi designing a continuous stirred tank, it is generally assumed that the contents are perfectly mixed so that the outlet flow has the same composition as the slurry inside the tank. However, in practice, these tanks are not perfectly mixed and non-ideal flow conditions occur, such as short-circuiting and stagnant zones. Short-circuiting results in low extractions because a significant proportion of the product has been in the reactor for less than the average residence time. To increase overall extraction, the system residence time should be divided into a series of stirred tank reactors 29  (Levenspiel, 1972). The greatest increase in extraction occurs i f a second reactor is added, the second greatest occurs i f a third reactor is added and so on. Typically in chemical engineering applications, three or four stirred reactors in series are used.  There is a trade-off between  maximizing conversion and minimizing capital cost. Mixing in the stirred tanks can also be improved with the use of internal baffles, which prevent swirling, and an entrance weir, which can help reduce short-circuiting effects (Gormely and Branion, 1989).  The development of bench scale systems for the continuous biological leaching of sulphide ores began in the 1960's with the use of batch stirred reactors (Duncan et al., 1967; Duncan and McGoran, 1971). Then in 1971, Bruynesteyn and Duncan proposed a continuous stirred tank leaching process for bioleaching sulphide concentrates based on the results of batch tank leaching experiments. Gormely et al. (1975) demonstrated the feasibility of bioleaching a zinc sulphide concentrate in a bench scale single-stage continuous stirred tank reactor. Several years later, a larger single stage reactor was used for the continuous bioleaching of a chalcopyrite concentrate. However, significant short-circuiting was found to occur and system upsets due to mechanical problems with solids feeding occurred (McElroy and Bruynesteyn, 1978). Sanmugasunderam (1981) operated a two-stage bench scale continuous biological leaching system for leaching zinc sulphide concentrate. The effect of the second stage was to increase the ultimate zinc extraction and the rate of zinc extraction (Sanmugasunderam et al., 1986). A continuous biological leaching system with five stages of 5 L each was used to test six different copper sulphide concentrates, which were washed in acetone to eliminate flotation chemicals (Groudev, 1985). The highest final copper extractions were achieved with the secondary sulphide concentrates. Corral et al. (1993) commissioned a 1.6 L three-stage continuous reactor system for the bioleaching of sulphide minerals. Two chalcopyrite concentrates were tested, and the pulp density and residence time were varied. Data reported for several 30  continuous bioleaching experiments are given in Table 2.3.  Table 2.3 - Summary of Several Continuous Bioleaching Studies Reference:  Corral et al., 1993  Groudev, 1985  Groudev, 1985  Groudev, 1985  main copper mineral(s) in feed  CuFeS  CuFeS  CuS, Cu S, CuFeS  CuS, Cu S, CuFeS  2  2  2  2  2  2  total copper in feed (%)  26.9  21.2  21.3  35.0  number of stages  3 (1.6 L each)  5 (5 L each)  5 (5 L each)  5 (5 L each)  pulp density (%)  5  20  20  20  residence time (days)  4.0  4.3  2.7  5.7  [Cu] (g/L)  6.5  26  30  58  copper leach rate (mg/L/h)  not given  250  470  430  copper extraction (%)  43  61  71  83  out  Although the feasibility of continuous bioleaching of chalcopyrite concentrates was demonstrated, it was found that leach residues had to be reground and releached in order to achieve reasonable copper recoveries (McElroy and Bruynesteyn, 1978; Groudev, 1985; Briceno and Rossi, 1987). A two stage continuous bioleaching system for the silver catalysed bioleaching of chalcopyrite concentrate was run for about 35 days (Lawrence et al., 1984). They found that most of the leaching occurred in the first tank, and they concluded that short-circuiting occurred because copper recoveries were lower than had been achieved in batch reactor tests.  Continuous stirred tank biological leaching for refractory gold ores and concentrates has also been carried out successfully at the bench scale (Nagpal, 1993; Pinches et al., 1988).  31  However,  biooxidation of refractory gold ores and concentrates is further advanced than biological leaching for copper extraction. Many pilot plant projects have been successful (Lawrence, 1993; Adam et al., 1989; Hackl et al., 1989; Moore, 1989) and the process is also successful on a commercial scale.  A mathematical model is an equation or set of equations that describes the reactions occurring in a system. Mathematical models allow the prediction of system behaviour under many more conditions than it would be practical to test in the laboratory. This is particularly important for continuous bioleaching studies, which are time consuming and labour intensive. However, biological leaching systems cannot be described mathematically without some experimental data on leaching kinetics. Several mathematical models of continuous biological leaching processes have been developed (Petrova, 1985; Konishi and Asai, 1993; Nagpal, 1993). These models were all based on some form of Monod growth kinetics. Blancarte-Zurita (1988) used a shrinking particle model to describe biological leaching of sulphide mineral concentrates. A computer program was then written to solve this model for continuous stirred bioleach tanks in series. A n empirical equation relating the rate of leaching to the mineral particle size is a required input to the computer program.  2.7  Commercial Processes  Several excellent reviews of the commercial application of biological leaching processes have been published (Hiskey, 1994; Marchbank, 1993; Rossi, 1990; Adam et al., 1989). Commercial biological leaching operations  for extraction of copper from sulphide ores are mainly in-situ and dump  leaching. Copper extraction by heap leaching was carried out for many years at the Rio Tinto mine in Spain. However, heap leaching has only reached widespread commercial application in the past few years. Continuous stirred tank bioleaching for copper extraction is still at the laboratory stage. Presently, there are no commercial scale biological leaching processes for copper sulphide 32  concentrates. However, there are several commercial continuous stirred tank biooxidation plants for treatment of refractory gold ores and concentrates (Rhodes et al., 1995; Lawrence, 1993).  Copper extraction by in-situ leaching is carried out at the Old Reliable mine and the Inspiration mine in the United States, the Kosaka mine in Japan and the E l Teniente mine in Chile (Rossi, 1990, p.498). In-situ leaching for copper extraction is also carried out at the Blyavinsk mine in the USSR and the Avoca mine in Ireland (Karavaiko et a l , 1988, p.285-292). Biological copper dump leaching is carried out at the Kennecott Bingham Canyon (Malouf and Prater, 1961) and Esperanza mines in the United States, a mine at Cananea, Mexico (Rossi, 1990, p.499), the Gibraltar mine in Canada (Bruynesteyn et al., 1987) and the Nikolayev and Kounrad mines in the USSR (Karavaiko et al., 1988, p.277-285). The Phelps-Dodge Morenci mine in Arizona, U S A has been operating a copper sulphide biological heap for many years (Marchbank, 1993). Recently, a number of biological heap leaching operations have been developed for extraction of copper from sulphide ores. In Chile, the Lo Aguirre mine began bacterial heap leaching in 1988 and the Cerro Colorado and Quebrada Blanca mines opened in 1994 (Marchbank, 1993). Biological heap leaching of a copper ore will commence at the Zaldivar mine in Chile in the near future.  2.8  Summary  It is clear from the literature review that the amenability of individual ores and concentrates to biological leaching must be determined by experiment. This is particularly true for concentrates because higher metal contents and the presence of flotation chemicals can potentially inhibit bacterial growth. The results of previous studies suggested that it is necessary to carry out initial batch tests in order to ascertain the feasibility of biological leaching and to determine ultimate extraction values. Continuous biological leaching experiments were determined to be necessary to obtain scale-up data 33  for commercial processes and for determining the adaptability of the bacteria to a particular feed material over time. The literature review also highlighted the importance of using a mixed bacterial culture that is already adapted to environmental conditions at the site under consideration for operation of a biological leaching process. No reports were found in the literature concerning the continuous biological leaching of pure chalcocitic ores or concentrates, although there were some data for mixed chalcopyrite and chalcocite feed concentrates. Also, continuous biological leaching for copper extraction in a saline environment has apparently not yet been studied.  The present study has considered the feasibility of the continuous biological leaching a chalcocitic ore and concentrate in a saline environment. Based on the information given in the literature, it was determined that the study would be carried out in three phases, including two initial phases of batch tests followed by the continuous bioleaching experiment. A mixed bacterial culture obtained from the mine site was used. The continuous biological leaching system was designed based on systems reported in the literature. A three-stage system was selected to achieve a high conversion with a reasonable amount of work involved in system maintenance.  The optimal values of several key variables were obtained from the literature, such as pH, temperature, oxygen, carbon dioxide and nutrient concentrations.  It had been reported in the  literature that the bacteria could be adapted to growth in the presence of chloride, flotation chemicals and high copper concentrations.  Therefore, it was anticipated that it would be possible to  biologically leach the Zaldivar ore and concentrate under these conditions, and no attempt was made to wash the concentrate to remove flotation chemicals. The literature review revealed that pulp density and initial soluble iron concentration were important variables in biological leaching. Therefore, they were selected as key variables for this study. Stirring and aeration were determined 34  to be critical parameters. A n axial impeller was selected for this study in order to achieve adequate solids suspension and air dispersion with low shear rates.  There have been several attempts to mathematically model continuous bioleaching processes. Apparently no mathematical models have yet been applied to the continuous bioleaching of copper sulphide ores. Mathematical modelling was not undertaken in this study due to time constraints.  Continuous stirred tank biological leaching has been successful commercially for the biooxidation of refractory gold ores. Copper extraction is another potential application for this process. The success of continuous bioleaching for the treatment of refractory gold ores will likely be extended to copper extraction when an economically viable system is developed.  35  3.0  EXPERIMENTAL METHODS  3.1  Feed Materials  The two feed materials used for this study were a rhyolite sulphide ore from the Zaldivar mine and a flotation concentrate produced from the ore by Placer Dome Inc. (PDI). The mineralogy included mainly the sulphide minerals chalcocite (Cu S) and pyrite (FeS ), and the copper oxide brochantite 2  2  (Cu (S0 )(OH) ). Brochantite is soluble in very dilute acids (Mottana et al., 1978). The other main 4  4  6  iron containing mineral was hematite (Fe 0 ). Analyses of the feed materials are given in Table 3.1. 2  3  Table 3.1 - Characteristics of the Zaldivar Ore and Concentrate Feed  Cu  s-  ^80  (%)  (%)  (um)  (%)  (%)  Fe (%)  Ore  1.56  0.50  2.66  0.91  0.81  108  2.70  Concentrate  17.30  1.60  11.60  11.70  11.62  53  3.32  x  Cu  Acid  Soluble  2  x  s.g.  The feed materials were supplied by PDI in slurry form at pulp densities of 47% and 33 % for the ore and concentrate respectively. Portions of each slurry were filtered to obtain dry ore and concentrate for use in the shake flask and batch reactor experiments. The filtration residues were dried at ambient temperature, rolled and separated into representative sections. Dried samples of each feed material were ground in a rod mill to obtain various particle size distributions for testing in the shake flask experiments. Fresh ore and concentrate slurries were used as the feed materials for the continuous bioleaching systems.  PDI attempted to produce a flotation concentrate with a total copper content of about 30%, which  36  was the expected copper content of the concentrate to be produced at the Zaldivar mine site. Unfortunately, the highest total copper content attainable was 17.3%. The poor copper grade of the concentrate was likely attributable to the Zaldivar ore having been stored outside in drums for over one year. As the drums were not tightly sealed, the ore was exposed to air and water during this time and was likely partially oxidized. The chemicals used in the flotation process are given in Table 3.2 along with their concentrations.  Table 3.2 - Flotation Chemicals Used in the Production of the Zaldivar Concentrate  3.2  Product Name  Chemical Name  Concentration (ppm)  Dowfroth (R) 250-D Frother  polypropylene glycol methyl ethers  15-20  sodium sulphide  sodium sulphide  250  T-22  isopropyl ethylthionocarbamate (85%) methyl ethyl-dithiocarbamate (7%) isopropyl methyl-xanthate (2%) dimethyltrithiocarbonate (2%) unknown organic impurities (4%)  30  Aero (R) 350  potassium amyl xanthate  60-75  Bacterial Culture and Nutrient Media  The bacterial culture used was a mixed culture dominated by Thiobacillus ferrooxidans. The culture was originally obtained from the Zaldivar site for a previous study at the University of British Columbia. It had been acclimatized to the Zaldivar ore and a saline environment of 9 g/L chloride before being used in this study. The bacterial nutrient solution used in the shake flask and batch reactor experiments was the 9K solution developed by Silverman and Lundgren (1959) containing 9 g/L ferrous iron. Similar solutions containing lower concentrations of ferrous iron were used in  37  shake flask and batch reactor experiments to test the effect of the initial soluble iron concentration. The bacterial nutrient solution used in the continuous reactor experiments was the Silverman and Lundgren solution (1959) containing 3 g/L ferrous iron. It was expected that the Zaldivar plant water would have this soluble iron concentration. A 3 g/L chloride solution was used because the Zaldivar plant water was expected to have this chloride concentration. The required chloride concentration was achieved by adding the appropriate amount of a 211 g/L NaCl stock solution.  Stock bacterial cultures were grown in shake flasks on the Zaldivar ore at a pulp density of 10% in a 3 g/L chloride environment. They were maintained by successive transfer of 5 mL of slurry containing cells in the late exponential phase of growth. This is known as the serial transfer method. During the continuous reactor experiment, stock bacterial cultures were maintained on both the ore and concentrate in 2 L batch reactors. When the cultures reached the late exponential phase of growth, 500 mL of slurry was removed from the batch reactor and replaced with fresh slurry. These cultures were available for reinoculation of the continuous bioleaching systems in case washout conditions occurred.  3.3  Shake Flask Experiments  3.3.1  Apparatus and Procedures  The shake flask experiments were carried out in 250 mL bottom-baffled Erlenmeyer flasks. Each flask contained a specific amount of feed material (depending on the required pulp density), 70 mL of bacterial nutrient solution, 3 g/L chloride and 5 mL of bacterial inoculum from a stock culture. Sterile control experiments were also carried out in which 5 mL of a bactericide (2 g/L thymol in methanol) was substituted for the bacterial inoculum. The flasks were incubated in a rotary shaker 38  at a stirring speed of 250 R P M and a temperature of 35 °C.  The shake flasks were inoculated approximately twenty-four hours after starting the experiment. Time zero was considered to be the time of inoculation. Acid consuming reactions occurred during the twenty-four hours before inoculation and caused the pH to increase. Therefore, it was necessary to stabilize the pH in the optimal range for bacterial growth before adding the bacterial inoculum. The pH was adjusted to pH 2 by adding 6M H S 0 . Next, 5 mL of bacterial inoculum was added 2  4  from a stock culture in the late exponential phase of growth. Bacteria on solids as well as in solution were transferred from the stock culture to the experimental flask. The shake flasks were maintained at a constant weight by addition of deionized water to compensate for evaporative losses. The p H was maintained at about pH 2 by adding 6 M H S 0 as required. When pH was an experimental 2  4  variable, it was controlled by adding either 6M H S 0 or 6M NaOH as required. 2  4  Dissolved copper and iron, as well as the pH and the Eh of the slurry were determined on a regular basis. In order to minimize loss of slurry, the pH and Eh probes were rinsed back into the shake flasks after each measurement. Upon completion of each experiment, the shake flask contents were filtered using the repulp method (refer to Section 3.7). The filtrate and wash solutions were analysed for copper and iron. The leach residue was analysed for copper, iron, total sulphur and sulphate sulphur.  3.3.2  Experimental Design  The variables tested in the shake flask experiments for the ore and concentrate are given in Table 3.3. Only one of each experiment was carried out. Sterile control experiments were also carried out for each variable tested. In addition, the bacterial culture was adapted to growth on the concentrate 39  using the serial transfer method and by gradually increasing the pulp density over a four month period.  Table 3.3 - Shake Flask Experimental Variables Variable  Ore  Concentrate  4 7 / 6 0 / 108  37/53  1.25/ 1.5/ 1.75/2.0  1.25/ 1.5 / 1.75 /2.0  5/10/15/20  2.5/5/7.5/10  soluble iron concentration (g/L)  0/3/6/9  0/3/6/9  temperature (°C)  30/35/40  30/35 /40  not tested  0/3  P particle size (nm) 80  pH pulp density (% w/v)  chloride concentration (g/L)  In tests where particle size was not a variable, particles sizes of P  80  108 um and P 53 u.m were used 8 0  for the ore and concentrate respectively. Unless soluble iron concentration was a variable, 9 K bacterial nutrient solution was used containing 9 g/L soluble iron.  3.4  Batch Reactor Experiments  3.4.1  Apparatus and Procedures  Upon completion of the shake flask experiments, larger scale batch reactor experiments were carried out in 2 L stirred tanks. The tanks and tank lids were manufactured of acrylic. A tank diameter to tank height ratio of approximately 0.7 was used. The tanks were vertically baffled with the baffle width being 0.1 times the tank diameter. The baffles were raised above the tank bottom to prevent stagnant zones from forming. Two impellers per tank were used. They were the four blade, forty-  40  five degree pitch down type with a diameter of 0.4 times the tank diameter. Titanium impellers and shafts were used due to the corrosive environment (chloride and ferric iron). A n attempt was made to use Teflon coated stainless steel impellers but this was not successful. The Teflon coating wore rapidly and flaked off due to abrasion from the solid particles. Variable speed mixers were used to provide constant stirring at approximately 800 R P M . This was the minimum speed required for complete particle suspension. At this stirring rate, the impeller tip speed was calculated to be approximately 2.4 m/s. Air was sparged into the tank below the bottom impeller. The air supply came from a compressor and was C0 -enriched to 1 % v/v C 0 using a C 0 gas cylinder and a gas 2  2  2  proportioner. A gas flowmeter was located at the inlet of each tank, and was used to regulate the C0 -enriched air flowrate at 0.25 L air/ L pulp/minute. The temperature in each tank was controlled 2  to 35 °C using an immersion heater.  Operation and maintenance of the batch reactors was similar to the procedure used for the shake flask experiments. The batch reactors were inoculated approximately twenty-four hours after starting the experiment. Time zero was considered to be the time of inoculation. The inoculum consisted of 130 mL of a stock shake flask culture in the late exponential growth phase. The batch reactors were maintained at a constant volume by addition of deionized water to compensate for evaporative losses. The p H was maintained at about pH 2 by addition of 6M H S 0 as required. 2  4  Dissolved copper and iron, as well as solution pH and Eh were determined on a regular basis. Upon completion of the experiment, the slurry was filtered using the repulp method (refer to Section 3.7). The filtrate and wash solutions were analysed for copper and iron. The leach residue was analysed for copper, iron, total sulphur and sulphate sulphur.  41  3.4.2  Experimental Design  The variables tested were pulp density and soluble iron concentration, as detailed in Table 3.4. Otherwise, conditions were the same as those used for the shake flask experiments. When soluble iron was a variable, the ore pulp density was fixed at 15%.  Table 3.4 - Batch Reactor Experimental Variables Variable  Ore  pulp density (%)  10/15/20  soluble iron concentration (g/L)  0/3/6/9  Concentrate •  5/7.5/10  not tested  Each experiment was carried out only once. Sterile control tests were not carried out for the batch reactor experiments.  3.5  Continuous Reactor Experiments  3.5.1  Apparatus  Two identical bench scale continuous reactor biological leaching systems were operated simultaneously for a period of four months; one system tested the ore and the other tested the concentrate. Each system consisted of three continuously stirred biological leaching reactors in series. The volume of the first stage reactor was 4 L, and the second and third stage reactor volumes were each 2 L . A schematic diagram of the system is given in Figure 3.1.  42  •a SH  n II  H ? r n II  re re II  a.  00  re rft a  3 3. * 35 K  f  O r e l S 1  re  2.  = -i  =  o rt  Tim  re o o  n  f to i  a  f>3  3  4^  o cr CD  o o  I'  c  w  13  o  o* 03 o'  t-  1  CD  fa o  cr  5' CTQ  t» CD  *0  3  The feed tanks were very similar to the batch reactor tanks described above except that they were not supplied with air and they were not heated. Continuous flow was simulated by pumping the feed intermittently two or three times per hour for less than 10 seconds each cycle, depending on the desired residence time. Intermittent pumping was used to maximize the feed flow velocity and ensure representative transferring of slurry.  The continuous biological leaching tanks were also very similar to the batch reactor tanks described above. In addition, they contained an entrance weir that forced the incoming feed to the bottom of the tank before being mixed in the bulk solution; this was to reduce short-circuiting of the feed solids. Gravity overflow and inflow tubes were attached to each tank at the appropriate height to achieve the desired tank volume. The overflow and inflow tubes between tanks were joined by flexible P V C tubing. A controller maintained the p H in the first stage reactor at about p H 2 by addition of 6M H S 0 . Acid was not added to the feed tank or the second and third bioleach stages. 2  4  The stirring rate was set at the minimum speed required for complete particle suspension. This was a range from about 600 R P M to about 1100 R P M . At these stirring rates, the impeller tip speeds were calculated to range from 2.3 to 4.2 m/s in the first stage bioleach reactor, and from 1.8 to 3.3 m/s in the second and third stage bioleach reactors.  Dissolved oxygen was measured using a dissolved oxygen meter with salinity compensation. A programmable timer was used to provide intermittent feed pumping. A pH probe was placed in each of the first stage reactors for continuous pH measurement. A l l pumps were variable speed peristaltic pumps. Tygon tubing was used to transfer acid, and silicone tubing was used to transfer the feed slurry. As in the batch reactor experiments, the impellers were the four blade, forty-five degree pitch down type. In the feed tanks, where the environment was not as corrosive as in the biological 44  leaching tanks, 316 stainless steel impellers were used. Due to the difficulty of obtaining titanium impellers with the correct diameter for the first stage reactor (2-7/8"), acrylic impellers were used. Titanium impellers were used in the second and third stage bioleach reactors.  3.5.2  System Operation  The system was commissioned in batch mode. When the Eh in all three tanks reached 800 mV (Standard Hydrogen Electrode reference, SHE), the feed pump was turned on to begin continuous mode operation. The dissolved oxygen concentration, p H and Eh in each bioleach tank were measured daily. Solution samples were taken daily to analyse for copper and iron. Product slurry generation was recorded daily in terms of volume and weight. The volume of sulphuric acid added to the first bioleach stage was recorded daily.  The criterion for determining when the system had reached steady state was a constant dissolved copper concentration for three consecutive days. On three different days at steady state, a slurry sample of approximately 100 mL was taken from each of the three bioleach tanks. At least one 100 mL slurry sample was taken from the feed tank during each steady state period. These samples were filtered using the repulp method (refer to Section 3.7). The leach residues were analysed for copper, iron, total sulphur and sulphate sulphur. The results were used to determine average percent copper extraction and average percent sulphate formation at that steady state condition. Next, one variable was changed and the process was repeated.  3.5.3  Experimental Design  Four steady state conditions were tested, as given in Table 3.5.  45  Table 3.5 - Design of the Continuous Biological Leaching Experiments Variables  Test 1  Test 2  Test 3  Test 4  pulp density (%)  10  10  10  20  flowrate (L/day)  1  2  3  3  total residence time (days)  8  4  2.67  2.67  pulp density (%)  5  5  5  10  flowrate (L/day)  1  2  3  3  total residence time (days)  8  4  2.67  2.67  Zaldivar Ore  Zaldivar Concentrate  3.6  pH and Eh Measurements  Gel-filled combination pH probes with a silver/silver chloride reference electrode were used to measure pH. A flat-bottomed pH probe was used for the shake flask experiments because this type of probe minimized adherence of slurry to the probe.  Eh was measured using a platinum redox electrode combined with a silver/silver chloride reference electrode. A constant value of 199 m V was added to all measured Eh values. This was the redox potential developed by the reference electrode relative to the standard hydrogen electrode at a temperature of 25 °C.  3.7  Repulp Filtration Method  Slurries were filtered using the repulp method. This method involved filtering the solids twice. After the first filtration, the filtrate volume was recorded and a sample of the filtrate solution was taken. Next the solids were washed in a volume of deionized water that was about four times the 46  filtrate volume. The purpose of this repulping step was to wash dissolved copper containing solution out of the first filter cake. The solids were suspended in this solution by stirring for about 10 minutes before being filtered again. After the second filtration, the wash volume was recorded and the wash solution was sampled. The filtrate and wash solutions were analysed for copper and iron concentrations. The final filtration residue was dried at ambient temperature for two or three days. Since most sulphide minerals do not oxidize appreciably below about 100°C (Vaughan and Craig, 1978), drying at less than 100°C would likely not cause significant errors in the total sulphur results. However, to minimize the tendency for oxidation of sulphide sulphur or elemental sulphur to sulphur dioxide, ambient temperature was used.  3.8  Sample Analysis  Samples were analysed by personnel at either the Placer Dome Research Centre or International Plasma Laboratory (IPL) of Vancouver. Solution samples were analysed for copper and iron by atomic absorption spectrophotometry (AAS). The filtration residues were split into several parts for various analyses. Part of the residue was analysed for copper and iron by acid digestion and A A S analysis of the resulting solution. The rest of the residue was analysed for percent total sulphur and percent sulphate sulphur. Total sulphur was determined by the Leco method. Sulphate sulphur was determined by boiling in 15% HC1 to selectively dissolve sulphate, followed by gravimetric determination of sulphur as BaS0 . The difference between percent total sulphur and percent 4  sulphate sulphur included both sulphide sulphur and elemental sulphur. This difference was used to calculate percent sulphate formation.  47  4.0  RESULTS AND DISCUSSION  The results of this study are presented in the following order: shake flask experimental results, batch reactor experimental results and continuous reactor experimental results.  The acid  consumption results for all three phases are presented and discussed in a separate section.  4.1  Shake Flask Experiments  4.1.1  Comparison of Ore and Concentrate Bioleaching  The results of the shake flask experiments showed that the bacterial lag phase was much longer with the Zaldivar concentrate than with the ore, particularly in the presence of chloride. This is illustrated in Figure 4.1, which gives the Eh as a function of time for both the ore and the concentrate in 3 g/L chloride media, and for the concentrate in 0 g/L chloride media. A significant effect of chloride on bacterial growth was not observed with the ore. These results suggest that the higher copper content of the concentrate and the presence of flotation chemicals had an inhibitory effect on bacterial growth. This effect was magnified in the presence of 3 g/L chloride. As will be shown, this inhibition of bacterial growth was overcome in the concentrate continuous bioleaching experiment. The stock bacterial cultures used in this study were previously adapted to growth on the Zaldivar ore. However, they had not previously been grown on the Zaldivar concentrate. It was anticipated that the culture could also be adapted to the concentrate.  4.1.2  Bacterial Adaptation to the Concentrate  Figure 4.2 illustrates the effect of bacterial adaptation to the concentrate on the solution redox potential. The bacteria were adapted to growth on the concentrate in a 3 g/L chloride environment 48  900  21 Time (days) Ore - 3 g/L chloride  28  35  42  -•- Concentrate - 3 g/L chloride  Concentrate - 0 g/L chloride Figure 4.1 -  Effect of Feed Material and Chloride Concentration on Bacterial Growth in Shake Flasks at a Pulp Density of 10% as Measured by Eh  900  400  14  Generation 1 Figure 4.2 -  21 Time (days) Generation 12  28  35  42  Sterile  Adaptation of Bacteria to the Zaldivar Concentrate at a Pulp Density of 7.5% in Shake Flask Tests in a 3 g/L Chloride Environment as Measured by Eh  49  using the serial transfer method and by gradually increasing the pulp density. Curves of Eh versus time for experiments in the first and twelfth generations of bacteria growing on the Zaldivar concentrate at a pulp density of 7.5% are shown. The Eh profile for a sterile control experiment at the same pulp density is also given. The twelfth generation was reached in about four months.  4.1.3  Typical Bioleaching Results  Figure 4.3 gives a typical copper extraction profile resulting from the biological leaching of the Zaldivar ore in a 3 g/L chloride environment. This experiment was carried out at a pulp density of 10%.  The results of a sterile control experiment are also given. Profiles of dissolved copper and  iron concentrations for the same experiment are given in Figure 4.4, and the corresponding solution Eh and p H profiles are given in Figure 4.5. Similar results for the biological leaching of the Zaldivar concentrate are given in Figure 4.6, Figure 4.7 and Figure 4.8. The apparent decrease in the final data point of most copper extraction and dissolved copper profiles was due to dilution of the filtrate during repulp filtration on the final day of the experiment. This apparent decrease was not observed for experiments that were terminated before copper extraction was complete.  In the shake flask experiments, the ultimate copper extractions were consistently very high for all variables tested. The average ultimate copper extractions were found to be 90% for the Zaldivar ore and 97% for the Zaldivar concentrate. As shown in Figures 4.3 and 4.6, a significant proportion of the copper extraction occurred before inoculation of the bioleach experiments. As will be discussed in Section 4.1.4, part of this non-biological copper extraction is attributable to the dissolution of the oxide copper mineral brochantite (Cu (S0 )(OH) ). Also, as will be discussed in Section 4.1.4, part 4  4  6  of the copper extraction that occurred before inoculation is attributable to chalcocite oxidation to covellite. 50  3  4  5  Time (days) Bioleach Figure 4.3 -  Sterile  Copper Extraction During the Shake Flask Biological Leaching of the Zaldivar Ore at a Pulp Density of 10% in a 3 g/L Chloride Environment  10  -•- Copper Figure 4.4 -  -•- Iron  Copper and Iron Concentrations in Solution During the Shake Flask Biological Leaching of the Zaldivar Ore at a Pulp Density of 10% in a 3 g/L Chloride Environment  51  2.25  X o.  1.75  400  pH  Eh Figure 4.5 -  Solution Eh and pH During the Shake Flask Biological Leaching of the Zaldivar Ore at a Pulp Density of 10% in a 3 g/L Chloride Environment  21 Time (days) Bioleach Figure 4.6 -  1.25  4 5 Time (days)  28  42  Sterile  Copper Extraction During the Shake Flask Biological Leaching of the Zaldivar Concentrate at a Pulp Density of 10% in a 3 g/L Chloride Environment  52  16 14 00  c u u  12 10  a o U  S 4  21 Time (days)  14  Copper Figure 4.7 -  28  35  42  Iron  Copper and Iron Concentrations in Solution During the Shake Flask Biological Leaching of the Zaldivar Concentrate at a Pulp Density of 10% in a 3 g/L Chloride Environment  900 800 700  2.25  r  X a.  >  a, 600  1.75  JS  500 400  14  21 Time (days)  35  1.25 42  pH  Eh  Figure 4.8 -  28  Solution Eh and pH During the Shake Flask Biological Leaching of the Zaldivar Concentrate at a Pulp Density of 10% in a 3 g/L Chloride Environment  53  The dissolved iron profile shown in Figure 4.7 is of particular interest. During the third week of the experiment, a significant decrease in dissolved iron concentration was observed, which coincided with an increase in dissolved copper concentration. The decrease in dissolved iron concentration was most likely due to jarosite precipitation. Once the dissolved copper concentration had reached its ultimate value, the dissolved iron concentration began to increase again. This phenomenon is also illustrated very clearly in Figure 4.9, which gives dissolved copper and iron profiles for the biological leaching of the Zaldivar concentrate in solutions with no added chloride. The increase in dissolved iron concentration was most likely related to pyrite oxidation. The dissolved iron profiles suggest that the majority of pyrite was oxidized after most of the copper leaching had occurred. This confirms that chalcocite and covellite leach preferentially to pyrite.  Figure 4.9 -  Copper and Iron Concentrations in Solution During the Shake Flask Biological Leaching of the Zaldivar Concentrate at a Pulp Density of 10% in a 0 g/L Chloride Environment  54  The shake flask experiments were terminated when copper leaching was complete, not necessarily when the maximum sulfate formation was reached. Therefore, most experiments resulted in very high copper extraction but much lower sulfate formation. As discussed above, incomplete sulfate formation was partly attributable to incomplete oxidation of pyrite. As will be discussed later, another likely factor was incomplete oxidation of elemental sulphur.  4.1.4  Sterile Control Experiments  The average copper extractions for the sterile control experiments were found to be 77% for the ore and 68% for the concentrate. Part of the copper extraction can be attributed to the dissolution of the oxide copper mineral brochantite:  Cu (S0 )(OH) + 3H S0 4  4  6  2  4  4CuS0 + 6H 0 4  2  (13)  However, based on the ratio of acid soluble copper to total copper in each of the feed materials, brochantite dissolution can only account for 32% and 9% of copper extraction from the ore and concentrate respectively. Therefore, some chemical oxidation of chalcocite must have occurred in the sterile control experiments. Chalcocite oxidation to covellite can account for a further 34% and 46% of the copper extraction observed in sterile control experiments on the ore andconcentrate respectively:  Cu S + 0.5O + H S0 2  2  2  CuS0 + CuS + H Q  4  4  2  (3)  Chalcocite oxidation occurs spontaneously but the kinetics are slow at ambient temperature and pressure. However, over the long leach times of these experiments, it was expected that some 55  chemical oxidation of chalcocite would occur. In order to maintain a pH of 2, it was only necessary to add acid to the biological leaching shake flasks at the beginning of each experiment. However, it was necessary to add acid to the sterile control shake flasks throughout the experiment. Since chalcocite oxidation is acid consuming, as shown in Equation 3, this is further evidence that chalcocite oxidation occurred in the sterile control experiments.  The remaining 11% and 13% of the copper extractions that were observed in the sterile control experiments on the ore and concentrate respectively, can be attributed to covellite oxidation:  CuS + 20  -» CuSO  2  (4)  A  However, less than 5% sulphate formation was found to occur in the ore sterile control experiments, and no sulphate formation was found to occur in the concentrate sterile control experiments. These results suggest that the covellite-sulphide was not oxidized to sulphate but to elemental sulphur, as shown in Equation 8:  CuS + Fe (S0 \ 2  -  4  CuS0  4  + S° + 2FeS0  4  (8)  In order for the above reaction to occur, there must have been some ferric iron in solution in the sterile control flasks. Some ferrous iron present in the bacterial nutrient media was likely oxidized to ferric iron as follows:  2FeS0  4  + 0.5O + H S0 2  2  -  4  56  Fe (S0 \ 2  4  +H0 2  (6)  This reaction is catalysed by bacteria but it also occurs chemically at much lower rates. Ferric iron must have been consumed as quickly as it was produced because no Eh increase was observed.  4.1.5  Effect of Particle Size  The effect of ore particle size on sulphate formation after five days of biological leaching is given in Figure 4.10. Particle size had no noticeable effect on ultimate copper extraction, copper leach rate or acid consumption. However, the results showed that for both the ore and concentrate, as the particle size increased, the sulphate formation at the end of the experiment was lower. As copper extractions were not affected by particle size, and were over 95% in these experiments, the incomplete sulphate formation was likely due to incomplete oxidation of pyrite or elemental sulphur. Sulphate formation for the two concentrate particle sizes tested, which were P P  80  80  37 microns and  53 microns, were 87% and 74% respectively after eleven days of biological leaching.  100  ,  QI  20  Figure 4.10-  1  :  .  i  40  i  i  ,  i  ,  60 80 Particle Size (P80, microns)  i  100  ,  :  I  120  Effect of Ore Particle Size on Sulphate Formation After 5 Days of Shake Flask Biological Leaching in a 3 g/L Chloride Environment  57  4.1.6  Effect of pH  There was no noticeable effect of the controlled pH value on ultimate copper extraction or copper leach rate. The controlled pH value was observed to effect the sulphate formation at the end of the experiment.  As the controlled pH value decreased, the  sulphate formation at the end of the  experiment decreased. For example, in an experiment on the concentrate at 10% pulp density, the pH was controlled to pH 2 and the sulphate formation was found to be 81% after 26 days of biological leaching. However, when the pH was controlled to pH 1.25, there was found to be no sulphate formation after 54 days of biological leaching. The lower sulphate formation was most likely related to lower oxidation of pyrite or elemental sulphur. Since pyrite oxidation is an acid generating reaction, it was inhibited in more acidic environments:  2FeS + 7.50 + H 0 2  2  2  -  Fe (SO \ + H S0 2  A  2  4  (5)  Another possibility is that the lower pH values inhibited the oxidation of elemental sulphur, which is also an acid generating reaction:  S° + 1.50  2  4.1.7  + H0 2  -  H S0 2  4  (9)  Effect of Pulp Density  Figure 4.11 and Figure 4.12 give the Eh profiles for the bioleaching of various pulp densities of the Zaldivar ore and the Zaldivar concentrate respectively in a 3 g/L chloride environment. The results from sterile control experiments are also given. The length of the bacterial lag phase was greater for 58  900  400  Figure 4.11-  4 5 Time (days) 5%  10%  20%  Sterile (20%)  15%  Effect of Ore Pulp Density on Biological Leaching in Shake Flask Tests in a 3 g/L Chloride Environment as Measured by Eh  900  400  14  2.5%  5%  21 Time (days)  28  7.5%  10%  35  42 Sterile (5%)  Figure 4.12 - Effect of Concentrate Pulp Density on Biological Leaching in Shake Flask Tests in a 3 g/L Chloride Environment as Measured by Eh  59  ore pulp densities over 10%, and increased with increasing pulp density for the concentrate. The increased lag time with increasing pulp density suggests that it took longer to establish a bacterial population at higher pulp densities. This may have been due to a limited supply of carbon dioxide at higher pulp densities in the shake flask experiments. Another reason may have been adherence of bacteria to solids. As the pulp density increased, the initial ratio of bacteria in solution to bacteria on solids may have decreased. Then, there would have been fewer bacteria in solution to oxidize ferrous to ferric iron, resulting in the longer observed bacterial lag phase at higher pulp densities.  There was no noticeable effect of pulp density on the ultimate copper extraction or on acid consumption. The copper leach rate was observed to be directly proportional to pulp density for both feed materials; it doubled with a doubling of the pulp density. The copper leach rate ranged from 4 to 18 mg/L/h for ore pulp densities ranging from 5 to 20%, and from 13 to 42 mg/L/h for concentrate pulp densities ranging from 2.5 to 10%. The sulphate formation at the end of the experiment was observed to decrease with increasing pulp density. This effect of pulp density on sulphate formation was likely due to incomplete oxidation of pyrite or elemental sulphur.  4.1.8  Effect of Initial Soluble Iron Concentration  Figure 4.13 shows the effect of initial soluble iron concentration on the bioleaching of the Zaldivar concentrate at a pulp density of 10%. The length of the lag phase appeared to be related to the initial soluble iron concentration, for both the Zaldivar ore and concentrate. Shorter lag phases were observed with higher initial soluble iron concentrations. Apparently, as shown in Figure 4.13, bacterial growth was inhibited during the experiment in which the initial soluble iron concentration was 6 g/L. This may have been due to insufficient transfer of bacteria when the shake flask experiment was inoculated. No significant effect of initial soluble iron concentration was observed 60  on the acid consumption, ultimate copper extraction, rate of copper leaching or sulphate formation.  0  7  14  21  28  35  42  49  56  T i m e (days) - • -  Og/L  3 g/L  — -  -«-  9 g/L  Sterile (9 g/L)  6 g/L  Figure 4.13- Effect of Initial Soluble Iron Concentration on Copper Extraction from the Zaldivar Concentrate at 10% Pulp Density in Shake Flask Tests in a 3 g/L Chloride Environment  The final solution pH was observed to be lower with higher initial soluble iron concentrations for both feed materials, as illustrated in Figure 4.14 for the biological leaching of the Zaldivar ore at 10%) pulp density.  This was likely due to increased jarosite precipitation at higher iron  concentrations. Since sulphuric acid is generated when jarosite precipitates, the more this reaction occurred, the more acidic the solution became:  3Fe (S0 \ 2  4  + \4H 0 2  ^  2H OFe (S0 ) (OH) 3  61  3  4  2  6  + 5H S0 2  4  (10)  4.1.9  Effect of Temperature  No effects of temperature on acid consumption, ultimate copper extraction or length of the bacterial lag phase were observed in the range tested of 30°C to 40°C. Both the copper leach rate and the sulphate formation at the end of the experiment were observed to increase with increasing temperature. The higher sulphate formation values were most likely due to higher rates of oxidation of pyrite or elemental sulphur. These effects were observed in the sterile control flasks, as well as in the bioleaching shake flask experiments. For example, the copper leach rate from the ore at 10% pulp density was found to increase from 9 to 23 mg/L/h in the temperature range 30 to 40°C. In the corresponding sterile control experiments, the rate was found to increase from 2 to 6 mg/L/h in the temperature range 3 0 to 40 ° C.  1 25 I 0  1  1  1  2  1  3  1  1  4  5  1  6  1  7  1  8  1  9  Time (days) •  Og/L  3g/L  9 g/L  Sterile (9 g/L)  6g/L  Figure 4.14 - Effect of Initial Soluble Iron Concentration on the Solution p H During the Shake Flask Biological Leaching of the Zaldivar Ore at 10% Pulp Density in a 3 g/L Chloride Environment  62  4.1.10 Summary The shake flask experiments demonstrated that the bacterial culture was successfully grown on both the Zaldivar ore and concentrate. The average copper extractions achieved in the shake flask tests were 90% and 97% for the ore and concentrate respectively. The higher copper extractions achieved for the concentrate were most likely due to the fact that the concentrate experiments were allowed to proceed longer than the ore experiments; the ore experiments were terminated before copper extraction was complete. The average copper leach rates were 12 mg/L/h for the ore and 50 mg/L/h for the concentrate, for pulp densities of 10%.  Average acid consumptions were 20 kg/t for the ore and 170 kg/t for the concentrate. The Eh ranged from initial values of about 500 mV to 900 mV (SHE) in active bacterial cultures. Sulphate formation, which was determined at the end of the experiment, was found to be significantly lower than the final copper extraction values achieved. Incomplete sulphate formation was thought to be due to incomplete oxidation of pyrite or elemental sulphur. Sample calculations for copper extraction, copper leach rate and sulphate formation are given in Appendix A . Selected shake flask experimental data is given in Appendix B .  4.2  Batch Reactor Experiments  4.2.1  Effect of Pulp Density  Figure 4.15 and Figure 4.16 give the copper extraction versus time for various pulp densities of the Zaldivar ore and the Zaldivar concentrate respectively in a 3 g/L chloride environment. As was observed in the shake flask experiments, the copper leach rate was directly proportional to the pulp density for both the ore and concentrate. There was an increase in the length of the lag phase at pulp 63  4 5 Time (days) 10% Figure 4.15-  15%  20%  Effect of Ore Pulp Density on Copper Extraction During Batch Reactor Tests in a 3 g/L Chloride Environment  21 28 Time (days) 5%  7.5%  49  10%  Figure 4.16 - Effect of Concentrate Pulp Density on Copper Extraction During Batch Reactor Tests in a 3 g/L Chloride Environment  64  densities over 10% for the ore, and with increasing pulp density for the concentrate, as was observed in the shake flask experiments. Pulp density had no noticeable effect on ultimate copper extraction, sulphate formation at the end of the experiment, or acid consumption. The increased lag phase at higher pulp densities may have been due to limited availability of carbon dioxide, which is necessary for bacterial growth. In the case of the concentrate, longer lag phases may have been partly due to the inhibitory effects of higher dissolved copper and higher concentrations of flotation chemicals. The fact that there was no noticeable effect of pulp density on ultimate copper extraction suggests that the bacteria were able to adapt to these conditions at higher concentrate pulp densities.  4.2.2  Effect of Leach Time on the Extent of Sulphate Formation  The extent of sulphate formation was found to increase with time after copper leaching was complete, all other experimental variables being equal. In duplicate batch reactor experiments on the Zaldivar concentrate, sulphate formations of 68% and 82% were achieved after 15 and 27 days respectively; corresponding copper extraction values were 95% and 97% respectively. As discussed in Section 4.1.3, incomplete sulfate formation is partly attributable to incomplete oxidation of pyrite. In addition, longer leaching times may have resulted in further oxidation of the elemental sulphur formed by the indirect biological leaching of covellite:  CuS + Fe (S0 ) 2  4  3  -»  CuSO + S° + 2FeS0 A  4  (8)  Figure 4.17 gives the dissolved copper and iron concentration profiles resulting from the duplicate batch reactor experiments carried out on Zaldivar concentrate. This graph clearly illustrates the increase in dissolved iron concentration between day 15 and day 27 of the longer experiment, which was likely due to pyrite oxidation. After approximately 15 days, the bacteria had oxidized all the 65  chalcocite and were left with only pyrite and elemental sulphur as a sources of energy.  0I 0  Figure 4.17-  4.2.3  i  i  i  i  i  i  i  7  i  i  i  •  i  i  i  i  i  i  i  14 Time (days)  i  i  i  •  21  -•-  Copper (15 days) - • - Copper (27 days)  -A-  Iron (15 days)  i  i  i  i  i  I  28  - • - Iron (27 days)  Copper and Iron Concentrations in Solution in Duplicate Batch Reactor Tests at a Pulp Density of 5% in a 3 g/L Chloride Environment  Effect of Initial Soluble Iron Concentration  Figure 4.18 gives copper extraction versus time for the bioleaching of the Zaldivar ore at 15% pulp density with several initial soluble iron concentrations. As shown in Figure 4.18, the initial soluble iron concentration had no observable effect on the copper leach rate or the ultimate copper extraction. There was also no noticeable effect of initial soluble iron concentration on sulphate formation at the end of the experiment, or on acid consumption.  Figure 4.19 shows the effect of initial soluble iron concentration on the pH profiles for batch reactor bioleaching experiments on the ore at 15% pulp density. As was observed in the shake flask 66  experiments, higher initial soluble iron concentrations resulted in lower final p H values in the batch reactors. This was likely due to increased jarosite precipitation and sulphuric acid generation when the soluble iron concentration was higher, as shown in Equation 10:  3Fe (S0 ) + UH 0 2  4.2.4  4  3  2  *=? 2H OFe (S0 ) (OH) 3  3  4  2  6  + 5H S0 2  4  (10)  Summary  The average percent copper extractions achieved in the batch reactor experiments were 90% and 97% for the ore and concentrate respectively. The higher copper extractions achieved for the concentrate were most likely due to the fact that the concentrate experiments were allowed to proceed longer than the ore experiments; the ore experiments were terminated before copper extraction was complete. The average copper leach rates were found to be 7 mg/L/h for the ore and 55 mg/L/h for the concentrate, when the pulp density was 10%. The average acid consumptions were 20 kg/t and 160 kg/t for the ore and concentrate respectively. The batch reactor results verified the shake flask results concerning the effects of pulp density and initial soluble iron concentration. The batch stirred tank reactor experimental method was developed successfully for use in the continuous experiments. Sample calculations for copper extraction, copper leach rate and sulphate formation are given in Appendix A. Selected batch reactor data are given in Appendix C.  67  100  14 Time (days) 0 g/L Fe Figure 4.18-  3 g/L Fe  21  6 g/L Fe  Effect of Initial Soluble Iron Concentration on Copper Extraction from the Zaldivar Ore at 15% Pulp Density During Batch Reactor Tests in a 3 g/L Chloride Environment  14 Time (days) 0 g/L Fe Figure 4.19-  28  3 g/L Fe  28  6g/LFe  Effect of Initial Soluble Iron Concentration on pH During the Batch Reactor Biological Leaching of the Zaldivar Ore at 15% Pulp Density in a 3 g/L Chloride Environment  68  4.3  Continuous Reactor Experiments  4.3.1  General Results and System Operation  A graph of soluble copper concentration versus time for the feed tank and the three stages of biological leaching for the ore run is given in Figure 4.20. A similar graph for the concentrate run is given in Figure 4.21. For approximately the first ten days of each experiment, the bacteria were grown in batch mode. After commencement of continuous mode operation, the copper level rose initially, partly because of excessive water loss due to evaporation. Water addition to compensate for evaporative losses was begun on approximately the twentieth day of each experiment. The amount of water added to the bioleach tanks was 40 mL water/L slurry for the ore run and 50 mL water/L slurry for the concentrate run. The evaporative losses were higher for the concentrate run because of the lower pulp density.  Foam was observed to form in all the bioleach tanks and solids were observed to coat the tank lids and tank walls above the liquid level. These solids were washed back into the slurry on a daily basis during the addition of water to compensate for evaporative losses. In addition to foaming, the accumulation of solids at the slurry surface may have been partly due to the presence of flotation chemicals in the concentrate slurry. In that case, the solids coating the tank lids and tank walls may have been sulphide-rich. As will be discussed in Section 4.3.4, this accumulation of sulphide-rich solids at the slurry surface was likely the cause of an accelerated short-circuiting effect observed in the concentrate continuous bioleaching system.  69  Time (days) -*»- Feed —*— Tank 1—•— Tank 2 — T a n k 3 Figure 4.20 - Copper Concentration versus Time for the Ore Continuous Biological Leaching Run in a 3 g/L Chloride Environment  16  0  7  14  21  28  35  42  49  56  63  70  77  84  91  98  105  112  119  Time (days) —Feed  —*— Tank 1—•— Tank 2—»— Tank 3  Figure 4.21 - Copper Concentration versus Time for the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment  70  During Test 4 of the ore run, mechanical failure caused simultaneous interruption of both the air supply and the acid supply to the bioleaching system. This resulted in a significant decrease in the activity of the bacterial culture. The feed pumps were turned off and the ore system was run in batch mode for several days. The bacterial culture recovered but unfortunately time constraints prevented the system from reaching steady state before completion of the experiment. However, as shown in Figure 4.20, before the mechanical failure occurred, a doubling of the pulp density resulted in a doubling of the copper concentration in the bioleaching tanks. This indicated that the system would likely have successfully reached steady state in the absence of mechanical problems.  It was difficult to continuously achieve the target flowrates for each test during the continuous bioleaching experiments, because of solid feeding problems. The feed tube often plugged during the night, particularly with the ore feed slurry since the P particle size was larger than that of the 80  concentrate feed slurry. McElroy and Bruynesteyn (1978) also reported difficulties with solids feeding in their continuous bioleaching system. The actual flowrate profiles for the ore and concentrate runs are given in Figure 4.22 and Figure 4.23 respectively.  Acrylic impellers were used in the first stage of both continuous bioleaching systems because titanium impellers of the correct diameter were difficult to obtain. Titanium impellers were used in the subsequent stages of the continuous bioleaching system. The blades of the acrylic impellers used in the ore system wore away over time because the P particle size of the ore was about double that 80  of the concentrate, making the ore more abrasive than the concentrate. In fact, the impeller blades used in the concentrate system wore only slightly during the four month experiment, while the ore blades wore away completely and required replacement. This decreased the amount of solids suspension in the first stage of the ore bioleaching system and reduced the amount of solids flowing 71  0  7  14  21  28  35  42  49  56  63  70  77  84  91  98  105 112 119  Time (days) —•—  Actual Flowrate  Target Flowrate  Figure 4.22 - Actual Flowrate for the Ore Continuous Biological Leaching Run in a 3 g/L Chloride Environment  Test 3  4> "cd  Test 4  3  u  o  1  2 Test 1  c o u o U  Batch Mode _l  7  14  21  28  35  42  49  I  I  I  I  I  l_  56  63  70  77  84  91  98  105 112  119  Time (days) Actual Flowrate  Target Flowrate  Figure 4.23 - Actual Flowrate for the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment  72  from the first to the second ore bioleach stage. Acrylic was found to be an unsatisfactory material for production of continuous bioleaching system impellers. Titanium impellers are recommended for future use.  4.3.2  Results at Steady State  Summaries of the continuous biological leaching results for the Zaldivar ore and concentrate are given in Table 4.1. For each steady state condition, the table gives the pulp density, total system residence time, redox potential in the third bioleach stage, copper leach rate in the first bioleach stage, cumulative sulphate formation in the third bioleach stage and cumulative copper extraction in the third bioleach stage. As mentioned above, due to mechanical failure near the end of the ore run, steady state conditions were not achieved for Test 4. Therefore, no results are given in Table 4.1 for continuous biological leaching of the ore at a pulp density of 20% and a total residence time of 2% days. The values for copper extraction and sulphate formation given in Table 4.1 are average values for three slurry samples taken during the steady state period. Three samples allowed confirmation that steady state had been achieved. The values for Eh and copper leach rate given in Table 4.1 are averages achieved during each steady state period. Sample calculations for copper extraction, copper leach rate and sulphate formation are given in Appendix A. Daily monitoring data for both the ore and concentrate continuous bioleaching systems are given in Appendix D .  The amounts of sulphuric acid required to maintain the pH below 2 in the first stage of continuous bioleaching were found to be approximately 50 kg per tonne of ore and 200 kg per tonne of concentrate. The net sulphuric acid requirements on an industrial scale would be lower than these values because some sulphuric acid would be recycled from the S X - E W circuit in an industrial process. This is discussed further in Section 4.4. 73  Table 4.1- Continuous Biological Leaching Steady State Results Variables  Test 1  Test 2  Test 3  Test 4  pulp density (%)  10  10  10  20  total residence time (days)  8  4  2%  2%  820  800  800  no results  stage 1 copper leach rate (mg/L/h)  6  10  16  no results  cumulative sulphate formation (%)  63  47  50  no results  cumulative copper extraction (%)  84  83  82  no results  pulp density (%)  5  5  5  10  total residence time (days)  8  4  2%  2%  stage 3 Eh (mV, SHE)  840  820  800  770  stage 1 copper leach rate (mg/L/h)  50  83  120  250  cumulative sulphate formation (%)  71  28  7  0  cumulative copper extraction (%)  91  89  87  83  Zaldivar Ore  stage 3 Eh (mV, SHE)  Zaldivar Concentrate  Profiles of the copper leach rate in the first bioleach stage are given in Figures 4.24 and 4.25 for the ore and concentrate runs respectively. As these figures and the data in Table 4.1 show, the copper leach rate was found to increase as the residence time decreased, while the overall copper extraction remained relatively constant. In order to maintain the same copper extraction with less residence time, the bacterial leach rate increased proportionately. Similarly, the leach rate was observed to double when the pulp density was doubled, while the copper extraction remained approximately constant. These results confirmed that chalcocite was readily amenable to oxidation by bacterial activity. In order to achieve copper extractions of 82% and 83% for the ore and concentrate respectively at 10% pulp density, a residence time of no more than 2% days was required.  74  Figure 4.24 - Copper Leach Rate versus Time in the First Bioleach Stage of the Ore Continuous Biological Leaching Run in a 3 g/L Chloride Environment  Figure 4.25 - Copper Leach Rate versus Time in the First Bioleach Stage of the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment  75  The stage 1 copper leach rates given in Table 4.1 were calculated based on the difference between the dissolved copper concentrations in the first bioleach stage and the feed tank. A check of these results was carried out as shown in Table 4.2. The stage 1 copper leach rate was used to calculate a theoretical copper extraction, which was then compared with the actual copper extraction observed in the first bioleach stage. This theoretical copper extraction is defined as the copper extraction that would have occurred at the stage 1 copper leach rate, for a given feed material, pulp density and residence time. A sample calculation of the theoretical copper extraction is given in Appendix A .  Table 4.2 - Check of Stage 1 Copper Leach Rates Test No. (Residence Time, Pulp Density)  Stage 1 Copper Leach Rate (mg/L/h)  Theoretical Stage 1 Copper Extraction (%)  Actual Stage 1 Copper Extraction (%)  Test 1 (8 days, 10%)  6  36  30  Test 2 (4 days, 10%)  10  30  47  Test 3 (2% days, 10%)  16  32  40  Test 1 (8 days, 5%)  50  55  49  Test 2 (4 days, 5%)  83  45  73  Test 3 (2% days, 5%)  120  44  72  Test 4 (2% days, 10%)  250  45  75  Zaldivar Ore  Zaldivar Concentrate  As shown in Table 4.2, the theoretical copper extraction was found to be in good agreement with the actual stage 1 copper extraction for the tests on the ore, and for Test 1 on the concentrate when the system residence time was 8 days. However, the theoretical copper extraction was found to be much  76  lower than the actual stage 1 copper extraction for subsequent tests on the concentrate. As will be discussed in Section 4.3.4, this was likely due to an accelerated short-circuiting effect observed in the first stage of the concentrate continuous bioleaching system when the residence time was 4 days or 2% days. The accelerated short-circuiting effect resulted in elevated actual copper extractions in the first bioleach stage. When this effect was not observed, the stage 1 copper leach rates were in good agreement with the actual copper extractions observed in the first bioleach stage.  Percent sulphate formation was found to be lower than percent copper extraction under all test conditions, as shown in Table 4.1. In addition, as the residence time decreased, copper extraction remained relatively constant while sulphate formation tended to decrease. As suggested by the results of the shake flask and batch reactor experiments, incomplete sulphate formation was likely due to incomplete oxidation of pyrite or elemental sulphur. In addition, according to the chemistry, it is possible to have the majority of the copper extracted and still have much of the sulphide unoxidized. Chalcocite oxidation involves the intermediate production of covellite:  Cu S + 0.5 0 2  2  + H S0 2  CuS0  4  + CuS + H 0  4  2  (3)  Covellite is then oxidized at a slower rate:  CuS + 20  CuS0  2  4  (4)  If most of the chalcocite is converted to covellite before any covellite oxidation occurs, then about half of the sulphide copper in a chalcocitic ore or concentrate is extracted before any sulphide 77  oxidation occurs. Based on the chemistry, it is possible to estimate the sulphate formation for a given copper extraction; a sample calculation is given in Appendix A . For example, the results of concentrate Test 4 given in Table 4.1, showed that there was 83% copper extraction but no sulphate formation. It was estimated based on the chemistry that in order to achieve 83% copper extraction, 22%o sulphate formation would occur, even i f there was no pyrite oxidation. Since no sulphate formation was observed, the covellite-sulphide must have been oxidized by the indirect mechanism to elemental sulphur, as shown in Equation 8:  CuS + Fe (S0 ) 2  4  3  -  CuS0  4  + S° + 2FeS0  4  (8)  Since Table 4.1 gives only the cumulative stage 3 copper extraction and sulphate formation, charts are included showing the copper extraction and sulphate formation by stage of the continuous biological leaching systems for each steady state condition tested. Figure 4.26, Figure 4.27 and Figure 4.28 give this data for the three steady state conditions tested in the ore run. Similarly, Figure 4.29, Figure 4.30, Figure 4.31 and Figure 4.32 are included for the concentrate run. In general, the largest proportion of the copper extraction occurred in the first stage of biological leaching. Correspondingly, the highest copper leach rate was observed in the first bioleach stage. For most test conditions, some sulphate formation occurred in the feed tank, and additional sulphate formation was observed in all three bioleach stages. A n interesting phenomenon was observed in the concentrate system under Test 2, 3 and 4 conditions. Lower values for copper extraction and sulphate formation were achieved in the second bioleach stage than in the first bioleach stage. This was most likely due to an accelerated short-circuiting effect in the first bioleach stage, as will be discussed in Section 4.3.4. 78  X  w  20  Stage Number Copper Extraction |  | Sulphate Formation  Figure 4.26 - Copper Extraction and Sulphate Formation for the Ore Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 8 Days and a Pulp Density of 10%  Stage Number Copper Extraction |  | Sulphate Formation  Figure 4.27 - Copper Extraction and Sulphate Formation for the Ore Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 4 Days and a Pulp Density of 10%  79  Stage Number Copper Extraction |  | Sulphate Formation  Figure 4.28 - Copper Extraction and Sulphate Formation for the Ore Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 2% Days and a Pulp Density of 10%  Stage Number Copper Extraction  [  | Sulphate Formation  Figure 4.29 - Copper Extraction and Sulphate Formation for the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 8 Days and a Pulp Density of 5%  80  X  w  20  Stage Number Copper Extraction |  | Sulphate Formation  Figure 4.30 - Copper Extraction and Sulphate Formation for the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 4 Days and a Pulp Density of 5%  100 0 s  80  o  u. o c  60  40  x w  20  Stage Number Copper Extraction |  | Sulphate Formation  Figure 4.31 - Copper Extraction and Sulphate Formation for the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 2% Days and a Pulp Density of 5%  81  100  a  80  a  8 o 60  40  x  w  20  Stage Number Copper Extraction |  | Sulphate Formation  Figure 4.32 - Copper Extraction and Sulphate Formation for the Concentrate Continuous Biological Leaching Run in a 3 g/L Chloride Environment with a Total Residence Time of 2% Days and a Pulp Density of 10%  4.3.3  Feed Tank Results  Copper extraction and sulphate formation were observed to occur in both the ore and concentrate feed tanks, particularly when the residence time was 8 days. Since the feed tank volume was 8 L and the total system volume was 8 L , the feed tank was refilled with fresh slurry every 8 days for a system residence time of 8 days, every 4 days for a system residence time of 4 days, and so on. The longer the residence time, the more copper extraction and sulphate formation occurred in the feed tank. In order to better understand the reactions that occurred in the feed tank, copper extraction and sulphate formation in the feed tank were plotted versus the time since the feed tank was refilled. As shown in Figure 4.33 and Figure 4.34, the copper extraction and sulphate formation in the ore feed tank increased as the age of the feed tank slurry increased. A similar correlation was observed in the concentrate feed tank, as shown in Figure 4.35 and Figure 4.36.  82  100  0  I  0  i  1  i  2  i  i  i  i  3 4 5 6 7 Age of Ore Feed Tank Slurry (days)  i  I  8  Figure 4.33 - Copper Extraction in the Ore Feed Tank versus the Age of the Feed Tank Slurry  Figure 4.34 - Sulphate Formation in the Ore Feed Tank versus the Age of the Feed Tank Slurry  83  100 80 60 x w S 40 o O  20 0  1  I  0  1  1  2  i  i  i  3 4 5 6 7 Age of Cone. Feed Tank Slurry (days)  i  I  8  Figure 4.35 - Copper Extraction in the Concentrate Feed Tank versus the Age of the Feed Tank Slurry  Figure 4.36 - Sulphate Formation in the Concentrate Feed Tank versus the Age of the Feed Tank Slurry  84  Part of the copper extraction observed in the feed tank is attributable to dissolution of the copper mineral brochantite (Cu (S0 )(OH) ): 4  4  6  Cu (S0 )(OH) 4  4  + 3H S0  6  2  -  4  4CuS0  4  + 6H 0  (13)  2  If it is assumed that all the acid soluble copper in the feed materials was brochantite and that all the brochantite dissolved in the feed tanks, then the resulting feed tank copper extractions would be 32% for the ore and 9% for the concentrate. The actual feed tank copper extractions were close to these values when the residence time was 4 days or 2% days, as shown in Figures 4.27 and 4.28 for the ore and Figures 4.30,4.31 and 4.32 for the concentrate. However, the feed tank copper extractions were significantly higher than the copper extractions based on dissolution of brochantite when the residence time was 8 days, as shown in Figure 4.26 for the ore and Figure 4.29 for the concentrate.  The proportion of the copper extraction greater than that expected solely based on brochantite dissolution was most likely due to the conversion of chalcocite to covellite. Chalcocite oxidation occurred either by direct or indirect oxidation, as shown in Equations 3 and 7 respectively:  Cu S + 0.5 0 2  2  + H S0 2  Cu S + Fe (S0 \ 2  2  4  4  -+  -  CuS0  4  + CuS + H 0  CuS + CuS0  4  2  (3)  + 2FeS0  (7)  4  It is unlikely that there was any significant covellite oxidation in the feed tank because covellite oxidation is kinetically slower than chalcocite oxidation (Beck, 1977). Even i f there was some covellite oxidation, it would still not explain the observed sulphate formation in the feed tanks. Since pyrite was the other main sulphide mineral, the sulfate formation was likely attributable to 85  pyrite oxidation:  2FeS + 7.50 + H 0 2  2  2  -  Fe (SO \ 2  A  + H S0 2  4  (5)  Figure 4.37 shows the proportion of overall sulphate formation that is attributable chalcocitesulphide and pyrite-sulphide by stage of the ore continuous bioleaching system for an 8 day residence time and a pulp density of 10%.  Figure 4.37 - Sulphate Formation and the Proportion of Sulphate Formation Attributable to Chalcocite-Sulphide and Pyrite-Sulphide by Stage of the Ore Continuous Bioleaching System for Test 1  The observed copper extraction and sulphate formation results, which occurred primarily when the residence time was 8 days, were likely due to the development of a bacterial population in the feed 86  tanks. Bacterial populations are associated with most ores, but they only become measurably active under certain conditions. Although no air was supplied to the feed tanks, some dissolved oxygen would have been available for biological activity due to the vigourous stirring required to maintain adequate solids suspension.  The average and range of pH and Eh observed in the ore and concentrate feed tanks are given in Table 4.3. Although pH 2.9 to 4.5 is out of the optimal range for T. ferrooxidans, these bacteria are still active in this p H range. There are also other mesophilic iron-oxidizing bacteria that are active in this pH range such as T. thiooxidans, L. ferrooxidans and the Metallogenium genus. It is possible that all of these bacteria were associated with the Zaldivar ore. Beck (1977) found that biologically catalysed chalcocite oxidation occurred at pH values less than about 4.6.  Table 4.3 - Average and Range of pH and Eh in the Ore and Concentrate Feed Tanks Feed Material Zaldivar Ore  Zaldivar Concentrate  •  pH  Eh (mV, SHE)  3.5 (2.9 to 3.8)  498 (465 to 544)  3.8 (3.6 to 4.5)  496 (446 to 563)  The pH and Eh in the feed tanks were both observed to increase with time after the tank was refilled with fresh feed slurry. One possible reason for the initially low feed tank pH was the acidity of the deionized water used for dilution of the feed materials. Also, acid may have been generated in the feed tanks by precipitation of ferric hydroxides; ferric sulphates may have been present in the feed materials due to partial oxidation during storage. Another possible source of acid may have been 87  the oxidation of other iron sulphide minerals such as marcasite. The acid in the feed tank was then available for the dissolution of brochantite:  Cu (S0 )(OH) 4  4  + 3H S0  6  2  4CuS0  4  4  + 6H 0  (13)  2  The observed pH increase with time in the feed tanks was most likely due to chalcocite oxidation, as shown in Equation 3 or Equation 7:  Cu S + 0.5O + H SO 2  2  2  Cu S + Fe (SO \ 2  2  A  -  A  -  CuS0  + CuS + H 0  4  2  CuS .+ CuS0  4  + 2FeS0  4  (3)  (7)  Ferric iron may have been available from the oxidation of pyrite, shown in Equation 5, or the oxidation of ferrous iron, shown in Equation 6:  2FeS + 7.50 + H 0 2  2FeS0  4  2  -  2  + 0.5O + H S0 2  2  Fe (S0 ) 2  -  4  4  + H S0  3  2  Fe (S0 \ 2  4  4  +H0 2  (5)  (6)  Pyrite oxidation also generates acid, which would then be available for brochantite dissolution (Equation 13), chalcocite oxidation (Equation 3) or ferrous iron oxidation (Equation 6). Jarosite or iron hydroxides may have precipitated in the feed tank to generate acid as shown below:  88  3Fe (S0 \ 2  + \4H 0  4  2H OFe (S0 ) (OH)  2  Fe {S0 \ 2  4  3  + 6H 0 2  3  4 2  + 5H S0  2Fe(OH)  6  3  2  + 3H S0 2  4  4  (10)  (11)  The observed increase in Eh with time may have been due to the generation of ferric iron by pyrite oxidation (Equation 5) or the oxidation of ferrous iron (Equation 6). Ferrous iron oxidation occurs spontaneously but the kinetics are much faster when a bacterial catalyst is available, as may have been the case in the feed tank when the residence time was 8 days.  4.3.4  Accelerated Short-Circuiting Effects  As illustrated in Figures 4.30, 4.31 and 4.32, a lower copper extraction was observed in the second bioleach stage of the concentrate continuous bioleaching system than in the first bioleach stage, when the residence time was 4 days or 2% days. This was thought to have been caused by an accelerated short-circuiting effect in the first bioleach stage, which was due to foaming and the presence of flotation chemicals in the concentrate feed. The flotation effect in the first bioleach stage created a sulphide-rich overflow to the second bioleach stage, reducing the observed copper extraction in the second bioleach stage. Pinches et al. (1976) reported a similar mineral flotation effect during the batch reactor biological leaching of a chalcopyrite concentrate.  Copper extractions in the third bioleach stage were higher than in the second bioleach stage, as shown in Figures 4.30,4.31 and 4.32. This suggests that the accelerated short-circuiting effect did not occur in bioleach stage 2. In other words, the flotation effect did not occur to the same extent, and consequently there was improved mixing of the slurry in the second bioleach stage. In the 89  second bioleach stage, there had been more time for flotation chemicals to decompose. Also, bacteria excrete wetting agents that tend to minimize the effect of flotation chemicals. Either of these factors could have reduced the flotation effect in the second bioleach stage.  Therefore, the accelerated short-circuiting effect observed in this study was likely not due to poor tank design but to a flotation effect caused by residual flotation chemicals in the concentrate. This led to a preferential transfer of a sulphide-rich slurry from the first to the second bioleach stage. These results suggest that four stages of bioleaching would probably be required to achieve adequate copper extraction in a full scale plant. Also, an outlet weir may help to retard the preferential transfer of a sulphide-rich froth from one stage to the next. In a larger scale system, transfer of slurry between stages could be accomplished by pumping from the mid-height of the tank, instead of by gravity overflow.  4.3.5  Ore versus Concentrate Continuous Biological Leaching  For a pulp density of 10% and a residence time of 2% days, the ore and concentrate steady state copper extractions were found to be 82% and 83% respectively. It is important to note, however, that some copper loss results from flotation. Therefore, the overall copper recovery for the concentrate was actually somewhat less than 83%, depending on the efficiency of the flotation process. A comparison of the costs of ore versus concentrate continuous biological leaching is critical to the selection of one feed material over the other for use in a larger scale operation. Continuous biological leaching of the ore would require larger bioleach tanks due to the presence of gangue minerals in the ore. However, as mentioned above, higher overall copper recoveries would likely be achieved in an ore system. One advantage of concentrate continuous biological leaching is a smaller plant size, since most of the gangue minerals are rejected during flotation. 90  Also, net acid consumption is typically lower for concentrates than for ores since concentrates contain a much lower proportion of gangue minerals. The main disadvantage of concentrate continuous biological leaching is copper loss during flotation.  4.3.6  Summary  A flowsheet of the continuous bioleaching system is given in Figure 4.38. A mass balance over the system is given to summarize the results achieved at steady state during the continuous bioleaching of the Zaldivar concentrate, for a residence time of 2% days and a pulp density of 10%. The oxygen consumption was determined to be 0.024 kg/day for a concentrate feed rate of 0.3 kg/day and an ultimate copper extraction of 83%. For the ore, the oxygen consumption was determined to be 0.003 kg/day for an ultimate copper extraction of 82% and a feed rate of 0.3 kg/day. This feed rate corresponds to a residence time of 2% days and a pulp density of 10%. Oxygen consuming reactions were the oxidation of chalcocite, covellite and pyrite as shown in Equations 3, 4 and 5 respectively:  Cu S + 0.5O + H S0 2  2  2  -  4  CuS0  CuS + 20  2  2  +H0 2  2  CuS0  2  2FeS + 1.50  + CuS + H 0  4  (4)  4  -  Fe {S0 \ 2  (3)  4  + H S0 2  4  (5)  The gross consumption of sulphuric acid was found to be 200 kg/t for continuous bioleaching of the concentrate, or 0.06 kg/day for a concentrate feed rate of 0.3 kg/day and according to the conditions given in Figure 4.38. The Zaldivar concentrate was found to be a net acid generator, as will be discussed in Section 4.4.  91  O  -J  o 0.043  P  W  ©  0.009  o v< cw  c  o  ©  o o  vo  o  oo  p  CD  P  P  3  ft  £. a  a  o  a  CB <-»•  t3 (re  CO  o o  00  <© CD  CD  ©  KJ  a ere p  concentrate  KJ  o  (kg/day) copper  i' ^5 o  *+>  (kg/day)  o  copper extraction  cs  O o a  (%) copper in solution  ft  (kg/day)  -- a o- o CD CD 3 a o  (kg/day)  a> o  copper in solids  oo  to O 0  a  P  n  1  P  P  Vi Vi  3  a P  ©  © o  a o  CA  CD  ft on ^+ CD P  oo ft CD  P  00  3 o  o ©  ON  o  O  KJ  The continuous biological leaching of both the Zaldivar chalcocite ore and concentrate in a 3 g/L chloride environment was successfully demonstrated.  The bacteria were able to adapt to  progressively lower system residence times and higher pulp densities. At the lowest residence time achieved of 2% days with a pulp density of 10%, the steady state copper extraction was 82% for the ore and 83% for the concentrate.  These copper extraction results are higher than the copper  extractions estimated for the heap leaching operation at the Zaldivar mine. However, they are lower than the copper extractions typically achieved at copper smelters. Addition of a fourth stage of continuous biological leaching may improve copper recovery, particularly in the case of the concentrate, for which an accelerated short-circuiting effect was observed. Also, heap leaching of the continuous bioleaching system residue, as shown in Figure 1.1, would increase the ultimate copper extraction. Finer grinding of the crusher ore fines or the flotation concentrate could potentially result in improved copper recovery for the same residence time. A mineralogical examination of the experimental bioleach residues would provide information as to the nature of the unleached copper minerals.  There appeared to be no difference in the time required for the ore and concentrate systems to reach steady state. Since bacterial washout was not experienced, shorter residence times and/or higher pulp densities than those tested may be possible. Also, copper leach rates are likely to improve as bacteria adapt to conditions in a continuous biological leaching system. In comparing the ore and concentrate runs, no inhibitory effect of copper content in the presence of chloride was observed in the case of the concentrate. Continuous biological leaching is worth further investigation as an alternative method of processing the ore crusher fines or the flotation concentrate at the copper mine of the Compania Minera Zaldivar in Chile. Other copper extraction operations with a secondary copper sulphide feed may also benefit from this continuous biological leaching system. 93  4.4  Acid Consumption  Gross acid consumption is defined as the amount of acid required to neutralize acid consuming minerals and dissolve copper minerals in the ore. Gross acid consumptions were calculated as the total amount of sulphuric acid added during an experiment to keep the pH at about 2, divided by the amount of substrate added. The average gross acid consumptions determined for each phase of the experimental work are given in Table 4.4.  Table 4.4 - Summary of Gross Acid Consumption Experiment  Zaldivar Ore (kgFLS0 /t)  Zaldivar Concentrate (kgH,S0 /t)  Shake Flask  20  170  Batch Reactor  20  160  Continuous Reactor  50  200  4  4  The acid consumption in continuous mode was found to be higher than in batch mode for both the ore and the concentrate. For the shake flask and batch reactor experiments, the p H was maintained between approximately pH 1.95 and p H 2. For the continuous reactor experiments, the gross acid consumptions were calculated as the amount of sulphuric acid added to the first stage of the continuous biological leaching system during a particular steady state period, divided by the amount of concentrate processed during that period. The average p H values for the ore and concentrate continuous reactor experiments were 1.78 and 1.83 respectively; it was difficult to control the p H to exactly 2. The lower pH accounts for the higher acid consumptions observed in the continuous reactor experiments as compared with the batch experiments. The difference between the acid consumptions observed in batch versus continuous experiments was much greater for the ore than 94  for the concentrate because a much lower proportion of the acid consumption was due to copper extraction in the case of the ore.  The pH is a very important factor in considering the acid  consumption results obtained for this continuous bioleaching experiment. In an industrial plant, the acid consumption would likely be lower than the results for this experiment since p H control would be more tightly controlled to pH 2, and acid would be recycled from SX-EW.  A breakdown of the calculated acid consumption for the ore and concentrate continuous reactor experiments by type of mineral is given in Table 4.5. Acid consuming gangue minerals in an ore include carbonates, such as calcite, and silicates. The gangue minerals in the Zaldivar ore were mainly quartz (Si0 ) and muscovite ( K A l ( A l S i ) O ( O H ) 2 ) . Other gangue minerals in the oxide 2  2  3  10  zone of the rhyolite ore were kaolinite (Al Si 0 (OH) ), rutile ( T i 0 ) , calcite ( C a C 0 ) , jarosite, 2  2  5  4  2  3  gypsum (CaS0 .2H 0) and hematite (Fe 0 ). 4  2  2  3  Table 4.5 - Percentage of Gross Acid Consumption by Type of Mineral Zaldivar Ore (kgH S0 /t)  Description  7  4  Zaldivar Concentrate (kgH S0 /t) 7  4  brochantite dissolution  5.8 (12%)  19 (10%)  chalcocite oxidation  8.2 (16%)  121 (61%)  ferrous ion oxidation  14.0 (28%)  23(11%)  gangue minerals  22.0 (44%)  37(18%)  TOTAL  50.0(100%)  200(100%)  1.78  1.83  28%  71%  average pH percent attributable to copper extraction r  95  The percentage of gross acid consumption that is attributable to copper extraction is important because it gives an indication of the amount of acid that can be recycled from the S X - E W circuit. Acid is regenerated during copper electro winning as follows:  CuS0  4  +H0 2  -  Cu + H SQ 2  4  + 0.5O  2  (14)  According to Equation 14, one mole of sulphuric acid is generated per mole of cathode copper produced.  Using the molecular weights of copper and sulphuric acid, it can be shown that  1.54 kilograms of sulphuric acid are generated per kilogram of cathode copper produced. Therefore, equations for net acid consumption can be written as follows:  ., . ., ^. i ca ( Cu extracted] net acid consumption = gross acid consumption - 1.54 V tonne ore )  (15)  net acid consumption = gross acid consumption - 1.54 j — C u extracted—j V tonne concentrate)  Using Equation 15, the net acid consumption for the Zaldivar ore was found to be about 30 kg/t. This was calculated based on a feed total copper content of 1.56% and a copper extraction of 82%. Using Equation 16, the Zaldivar concentrate was found to be a net acid generator of about 20 kg/t. This was calculated based on a feed total copper content of 17.3% and a copper extraction of 83%.  96  Several reactions in the system contributed to the overall acid consumption. Chalcocite oxidation by the direct mechanism (Equation 3), bacterial oxidation of ferrous to ferric iron (Equation 6) and brochantite dissolution (Equation 13) were all acid consuming reactions:  Cu S + 0.5 0  CuSO + CuS + H 0 2  (3)  2FeS0 + 0.5O + H S0  -  Fe (S0 \  + H0  (6)  Cu (S0 )(OH) + 3H S0  -  4CuS0 + 6H 0  (13)  2  + H S0  2  2  4  4  2  4  -  4  2  6  2  A  4  4  2  4  2  4  2  Acid generating reactions included pyrite oxidation by the direct mechanism (Equation 5), biooxidation of elemental sulphur (Equation 9) and jarosite precipitation (Equation 10):  2FeS + 1.50 2  2  + H0 2  S° + 1.50  2  3Fe (S0 ) + \4H Q 2  4  3  2  ^  -> Fe (S0 \ 2  + H0 2  -  + H S0  4  2  H S0 2  (9)  4  2H OFe (S0 ) (OH) 3  3  (5)  4  4  2  6  + 5H S0 2  4  (10)  Due to the complex chemistry, it was difficult to predict theoretically the acid requirements of the continuous biological leaching systems. However, for both the ore and concentrate runs, the acid consumption reached a constant value when the system was operating at steady state. Also, as expected, the acid consumption was approximately equal for tests in which the final copper extraction was approximately equal. 97  5J)  CONCLUSIONS AND RECOMMENDATIONS  5.1  Conclusions  1.  Shake flask and batch reactor experiments demonstrated that the bacteria were capable of oxidizing both the Zaldivar ore and concentrate in a 3 g/L chloride environment, and adapting to growth on these feed materials at various pulp densities.  2.  Under batch conditions, the higher copper content of the Zaldivar concentrate as compared to the ore appeared to have an inhibitory effect on bacterial growth, which was magnified in the presence of 3 g/L chloride. This inhibitory effect was overcome in the continuous biological leaching experiment.  3.  The Zaldivar ore and concentrate were found to be readily amenable to slow chemical oxidation, with 77% and 68% copper extractions respectively occurring in sterile control shake flask experiments.  4.  The residues from the biological leaching of both the ore and concentrate were thought to contain elemental sulphur, in addition to sulphide sulphur and sulphate sulphur. This suggests that covellite oxidation by the indirect ferric oxidation mechanism of biological leaching (Equation 8) was more important than covellite oxidation by the direct mechanism of biological leaching (Equation 4).  5.  Continuous biological leaching of a chalcocite ore and concentrate in a saline environment of 3 g/L chloride was successful. 98  6.  Continuous biological leaching is worth further investigation as a method of processing the ore crusher fines or the flotation concentrate at the Zaldivar mine in Chile. The solid residue from the continuous biological leaching process would be an excellent source of bacterial inoculum for biological heap leaching. Some of the copper remaining in the residue could be extracted by heap leaching. Higher copper extractions and leaching rates than those observed in this study are possible as bacteria adapt to system conditions over time. Other copper extraction operations with a secondary copper sulphide feed may also benefit from this continuous biological leaching system.  7.  The results of the ore continuous biological leaching experiment showed that the bacteria were able to adapt to progressively lower system residence times and higher pulp densities. The lowest residence time achieved was 2% days at a pulp density of 10%. The ultimate copper extraction at this steady state condition was 82% and the first stage copper leach rate was 16 mg/L/h.  8.  The results of the concentrate continuous biological leaching experiment showed that the bacteria were able to adapt to progressively lower system residence times and higher pulp densities. The lowest residence time achieved was 2% days and the highest pulp density achieved at this residence time was 10%. The ultimate copper extraction at this steady state condition was 83% and the first stage copper leach rate was 250 mg/L/h.  9.  Chalcocite was apparently biologically leached preferentially to pyrite from both the Zaldivar ore and concentrate. The main reasons for incomplete sulphate formation were thought to be incomplete oxidation of pyrite or elemental sulphur. 99  10.  A n accelerated short-circuiting effect was observed in the first bioleach stage of the concentrate continuous bioleaching system due to flotation of a sulphide-rich froth, and transfer of this froth from the first to the second bioleach stage. Four stages of biological leaching are probably required to achieve adequate copper extraction in a full scale plant.  Warning:  The results achieved in this continuous bioleaching experiment were partly due to leaching in the feed tank, particularly for the longest residence times tested. At a larger scale, it is not likely that leaching will occur in the feed tank because relatively low residence times will be used, and the feed material will be fresh and unoxidized.  5.2  Recommendations for Further Work at Zaldivar  The bench scale continuous biological leaching experiments carried out in this study are currently being continued at the Zaldivar site. The feed material being used is a flotation concentrate produced on-site. Following are several recommendations for these experiments at Zaldivar.  1.  The possibility of improved copper extraction could be investigated by studying the effect of finer grinding of the concentrate. In addition, a mineralogical examination of the bioleach residues would provide information as to the nature of the unleached copper minerals.  2.  Cost estimates of ore and concentrate continuous biological leaching could be used to compare the two systems.  3.  The amenability of the bacterial culture to biological leaching of the Zaldivar ore and concentrate without addition of the 9K nutrient media should be tested. The effects of 100  addition of the individual component nutrients in the 9K media could also be tested. The objective would be to determine what nutrients would be required in a larger scale continuous bioleaching plant.  A full analysis of the feed materials would provide  information about nutrients already present in the gangue minerals.  4.  The tests at the Zaldivar site are being conducted without C0 -enrichment of the air supply. 2  The results from the on-site tests should be compared with the results of this study to determine the effect of C0 -enrichment on the rate of bioleaching of the Zaldivar ore and 2  concentrate. It may not be necessary to provide carbon dioxide to a full scale bioleaching process. However, i f the lack of C0 -enrichment of the air supply is shown to have a 2  detrimental effect, a cheap source of CO should be found, such as carbonate minerals or 2  combustion gases.  5.3  General Recommendations  1.  Titanium was found to be the best material for the impellers and impeller shafts in stirred reactor biological leaching experiments carried out in a saline environment.  2.  Elemental sulphur analysis of bioleach residues would provide further information about the chemistry of bioleaching. The results of the elemental sulphur analysis could be used to evaluate the relative importance of the direct and indirect mechanisms of bioleaching.  3.  A mathematical model of the continuous bioleaching system used in this study could be developed and compared with the experimental results. Copper leach rates achieved in this study could be used as input to the model. 101  6J)  REFERENCES  Adam, K., Stefanakis, M . and Kontopoulos, A., Bacterial Oxidation of the Olympias Arsenical Pyrite Concentrate, In: Biotechnology in Minerals and Metal Processing, B.J. Scheiner, F . M . Doyle and S.K. Kawatra (Eds), Society of Mining Engineers, Inc., Cushing-Malloy, Inc., Ann Arbor, MI, U S A , 131-138,(1989) Agate, A . D . and Khinvasara, N.J., Bioleaching of Copper Ores and Concentrate of Malanjkhand Area, India, In: Workshop on Biotechnology for the Mining, Metal-Refining and Fossil Fuel Processing Industries, Biotechnol. Bioeng. Symp., No. 16, H.L. Ehrlich and D.S. 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Data:  = = = = = = = =  [C solutioiJ U  ^solution  [Cu ] V" inoc [Cu ] inoc  media  ^media £C-U  s a m  Mass  pj  e  V  s a m  p|  e  feed  \Cli / - ]  Copper Extraction = 100  so ttft  on  8.25 g/L 1.944 L 1.79 g/L 0.130 L 0.001 g/L 1.820 L 0.0583 g 99.7g VCu W  Vsolution  in  h  tnoc inoc  ]  J  [^^media^ ^media  yj cu  17 3 (——) Mass, ig) 100 K  /  m  A  (8.25)(1.944) - (1.79)(0.130) + (0.001)(1.820) + 0.0583 Copper Extraction, = 100 17 3 (99.7) 100 Copper Extraction, = 92%  Next, the copper calculated head must be determined. Data:  = = = =  Cu Cu Cu Mass feed out  inoc  media  16.391 g 0.233 g 0.002 g 99.7 g  out(s)  Cu  ~ Cu.(g)  Calculated Head = 100 feeM)  MaSS  109  +  Cu,(g)  v  sample  sample  Appendix A - Sample Calculations  16.391  0.233 + 0.002  Calculated Head =100  99.7  Calculated Head - 16.2%  Finally, the copper extraction based on copper in solution and calculated head can be determined.  Copper Extraction = Copper Extraction  x  Copper Extraction = 92% 2  17.3  Calculated Head  17.3 16.2  Copper Extraction., = 9 8 % Therefore, copper extraction based on copper in solution and calculated head was found to be 98%. Copper Extraction Based on Outputs This sample calculation is based on the results for batch reactor experiment BC11, which tested the bioleaching of the concentrate at a pulp density of 5%. Data:  = = = =  Cu,out, solids Cu,out Cu,-inoc Cu,media  0.414 g 16.391 g 0.233 g 0.002 g  C out, solids^) U  Copper Extraction = 100 out(g)  Cu  -  inoSS)  Cu  110  +  Cu,{g)  Appendix A - Sample Calculations  Copper Extraction = 100 1 -  0.414 16.391 - (0.233 + 0.002)  Copper Extraction = 97% Therefore, the copper extraction based on outputs was found to be 97%. Copper Leach Rate This sample calculation is based on results from the batch reactor experiment BC11 on 94/8/31. Data:  Rate =  Rate =  = 7.35 g/L = 6.20 g/L = 48h  [Cu], [Cu],, 5t  [Cu\ (g/L) - [ C u ] 5t (h)  M  (g/L)  1000 mg S  7.35 - 6.20 1000 48  Rate = 24 mg/L/h Therefore, the copper leach rate was determined to be 24 mg/L/h.  Ill  Appendix A - Sample Calculations  Sulphate Formation This sample calculation is based on the results for batch reactor experiment BC11, which tested the bioleaching of the concentrate at a pulp density of 5%. Data:  Solids (S -+S ) Solids Sulphide 2  0  out feed  = 2.038 g = 11.585 g  Solids (S ~ + S °) 2  out  Sulphate Formation - 100  Solids Sulphide^  Sulphate Formation = 100 1 -  (g) (g)  2.038 11.585  Sulphate Formation = 82% Therefore, the sulphate formation was determined to be 82%.  CONTINUOUS REACTOR EXPERIMENTS Copper Extraction This is an example of the calculated head method of determining percent copper extraction. The sample calculation uses data from a 100 ml slurry sample taken from the third concentrate bioleach stage on 94/12/5. This was during Test 1, when the residence time was eight days and the pulp density was 5%. Data:  Filtrate Wash Solids Total  Copper Extraction  94.5 ml 445 ml 4.38 g  =  8.35 g/L Cu 0.104 g/L Cu 1.95% Cu  100 * (1 - 0.085/0.920) 91% 112  0.789 g Cu 0.046 g C u 0.085 g Cu 0.920 g Cu  Appendix A - Sample Calculations  Therefore, the copper extraction was determined to be 91%. The average copper extraction for the three 100 ml slurry samples taken during this steady state period was also determined to be 91%.  Copper Leach Rate This is a sample calculation for the first concentrate bioleach stage on 94/12/5. This was during Test 1, when the residence time was eight days and the pulp density was 5%. Data:  [Cu] [Cu] ., V feed flowrate [Cu] , [Cu] .„ 8t  = = = = = = =  N  N  N  N  t  N  Rate -  [Cu] - [Cu] _ N  N  feed flowrate  Rate  6.65 - 1.27 4.0 0.824  {  6.65 g/L 1.27 g/L 4.0 L 0.824 L/day 6.65 g/L 6.55 g/L 23.4hr  [Cu]  - [Cu\ _  Nt  Nt  j6r 24  6.65 - 6.55 23.4 24  1000 24  1000 24  Rate = 50 mg/L/h Therefore, the copper leach rate for the first concentrate bioleach stage on 94/12/5 was determined to be 50 mg/L/h.  Check of Bioleach Stage 1 Copper Leach Rate This is a sample calculation for Test 1 of the concentrate continuous reactor bioleaching, when the overall residence time was 8 days and the pulp density was 5%. The overall stage 1 copper extraction based on solids analysis was found to be 85%. The copper extraction in the feed tank was found to be 36%. The actual copper extraction in bioleach stage 1 was the difference or 49%. Now, the theoretical copper extraction can be calculated based on the stage 1 copper leach rate.  113  Appendix A - Sample Calculations  Data:  Stage 1 Copper Leach Rate Stage 1 Residence Time Pulp Density Head Assay Concentrate Solids s.g. X  = 50 mg/L/h = 4 days = 50 g/Lpulp = 17.3% total copper = 3.32 = 0.985  ^solution I Lpulp  Copper Extraction = 100  Rate( L  ^  Cu  solution  Pulp Density(  h  ) Residence Time(days) (c^—) ( day 1000mgCu (  8,concentrate ) Head  Assay{-^—)(^-)  •"pulp  Copper Extraction  (50) (4) (24) (—L_) 1000 - 100 17 3 l ) 50 ( ^ ) ( 100 0.985  Copper Extraction  = 55%  & concentrate  *  solution  Therefore, the theoretical copper extraction was calculated to be 55%. Sulphate formation This sample calculation uses data from a 100 ml concentrate slurry sample taken from the third bioleach stage on 94/12/5. This was during Test l , when the residence time was eight days and the pulp density was 5%. The feed sulphide-sulphur assay was 11.62%. Feed Feed  5.00 g 5.00 g  17.3% Cu 11.62% S -  0.865 g Cu 0.581 g sulphide  Filtrate Wash Solids Total  94.5 ml 445 ml 4.38 g  8.35 g/L Cu 0.104 g/L Cu 1.95% Cu  0.789 0.046 0.085 0.920  Solids  4.38 g  3.45% (S- + S°)  0.151 g (S" + S°)  2  2  114  g g g g  Cu Cu Cu Cu 2  Appendix A - Sample Calculations  (S - + S°) ,.,(g) ^ 2  Sulphate Formation = 100 [1 -  (Sulphide^))  (  C  U  ]  actuals  ^"^)  feeM)  Cu  Sulphate Formation = 100 [1 -  0.151 0 920 (0.581) ( - ^ ^ ) 0.865  Sulphate Formation - 76% Therefore, the sulphate formation was determined to be 76%. The average of three 100 ml slurry samples analyzed during this steady state period was 71% sulphate formation. Overall Sulphate Formation Based on Chemistry This is a sample calculation to check the validity of the sulphate formation results. This calculation uses data from Test 4 of the concentrate continuous reactor bioleaching. The copper extraction based on mass balance analyses of the 100 mL slurry samples was found to be 83% and there was found to be no sulphate formation. Data:  Total copper assay Acid soluble copper assay Sulphide sulphur assay Total iron assay Copper molecular weight Sulphur molecular weight  =17.3% = 1.60% = 11.62% =11.6% = 63.546 g/mol = 32.06 g/mol  Assume the difference between total copper and acid soluble copper (15.7%) is chalcocite-copper. For a 100 g sample of dry concentrate feed: mass copper = 15.7 g moles copper = 15.7/63.546 = 0.247 moi moles chalcocite = 0.247/2 = 0.1235 moi moles sulphide as chalcocite =0.1235 moi mass sulphide as chalcocite = 0.1235 (32.06) = 3.96 g  115  Appendix A - Sample Calculations  total mass sulphide in 100 g percent of total sulphide as chalcocite  = 11.62 g = 100 (3.96/11.62) = 34%  From the chemistry: Cu S + 0.5O + H S0 2  2  2  4  CuS + CuSQ + H Q 4  (3)  2  CuSO  CuS + 2Q  (4)  A  2  Assume that all the chalcocite oxidizes before any covellite oxidation occurs. Therefore, half the total copper is extracted before any sulphide oxidation occurs. Then, to achieve overall 83% copper extraction, a further 33% copper extraction is required. This corresponds to 66%> covellite oxidation. If 66%) covellite oxidation occurs, then 66% of the chalcocite-sulphide must be oxidized to elemental sulphur or sulphate. As calculated above, the chalcocite-sulphide is 34% of the total sulphide in the feed. Assume that no pyrite oxidation occurs. Let X represent the overall sulphide oxidation. pyrite-sulphide  +  chalcocite-sulphide  (0%) (66%)  +  (66%) (34%)  X  =  =  total sulphide X(100%)  22%  Therefore, it is possible to have 83% copper extraction and only 22% total sulphide oxidation, assuming no pyrite oxidation occurs. It was found that there was no overall sulphate formation for Test 4 of the concentrate continuous bioleaching experiment. Therefore, the sulphide must have been oxidized to elemental sulphur.  116  Appendix B - Selected Shake Flask Data  Summary of Shake Flask Bioleach Experiment Expt. Name: R24 Expt Type: Bioleach - 9K Ore: Ore CI-Level: 3 gpl Fe(2+) Level: ~9 gpl Revision Date: April 16/95 fNPUTS: Solids Ore  Weight (g) Wt% Cu 8.02 1.56  Wt% Fe 2.66  Wt% S-2 0.81  WtCu  WtFe  WtS-2  (g)  (g)  (g)  0.125  0.213  0.009 0.000 NA NA  0.019 0.655 NA NA  0.134  0.887  0.065  ' Solutions Inoculum Media Acid 211 gplNaCl  Volume(mL) 5 70 0.25 1.726  System Volume (neglecting acid):  gpl Cu 1.72 0 NA NA  gpl Fe 3.75 . 9.35 NA NA  76.726  Total IN (g):  OUTPUTS: Solids Leach Residue  Weight (g) Wt% Cu 9.6300 0.109  Wt% Fe Wt% (S-2+S°) 6.840 0.11  0.065  WtCu  WtFe  Wt% (S-2+S°)  (g)  (g)  (g)  0.010  0.659  0.0009 0.0013 0.0013 0.0014 0.1103 0.0265  0.00850 0.00705 0.00418 0.00355 0.22950 0.02200  0.152  0.933  -0.019 -13.87%  -0.047 -5.29%  0.011  Solutions Volume(mL)  gpl Cu  gpl Fe  Sample # 1 2 3 4 Filtrate Wash  1 1 1 1 75 265  0.91 1.32 1.34 1.44 1.47 0.1000  8.5 7.05 4.18 3.55 3.06 0.083 Total OUT (g): Input - Output: % Lost of input:  Acid Addition Summary Volume 6M H2S04 added (ml): Acid Consumption (kg H2S04/tonne cone):  0.25 18.34  Copper Extraction Summary Copper Extraction based on Outputs (%): Calculated Head Copper (%):  92.7 1.79  Sulphate Formation Summary Sulphate Formation (%):  83.7  117  0.011  ^ 0 "2  = z 3 E_, C N ^  O C No o  o o o  O o o  o o o  o u-i o o C No o o ©  o o o  o o o  0 3  Tf  ^-^  T3  (/}  I  V D C N X)  EC "  in » n od ©  a. u-  .H  ON  3  o  u  a,  C N  m  00 m V D — •nO cn  Tf T f Tf m Tf Tf c^  >n O v m I O </"! vo C N< NV OO O >n V O oo oo 00  Tf Tf T f Tf m  ov oo O V  S3 a.  VI Tf  o o o  OV o O V— o cn i/S od  o  in o r~ o  —•  w  B Q  Q  t~  £  —I  Tf Tf  — f  w  Tf Tf Tf Tf Tf Tf  Vi  -a  gv ov O vO vO v — cn v V OO CJ ( D ej c C c c c 3 3 3 3 3  op  CO  ov  O v in  u  E  O v  T3  a.  CU CU  <  cd cd  N  o  Tf  ©  E ' ^—  3  1  <o  ext  X  ^i  in C N rC N C N 00  H  Tf  V OV D 00 o C NV O od cn vd V O vo r -  Tf  V OC No r~ m cn C N r~ o vd V D cn m C N  t~  O v o cn cn 00 vd O v r~ r - r~ oo 00  -J  W -J Cu  < X u  < w  —1  N  cd c  3  B — o © o m oo  m  o  C N O V O  1  3 <u ~ bu  3  Q  >  "8  E  o  3  _  3  1  U  3  •3  < w  .2  ^  O  o o >  -&> ^ o> o  G>  I !ai II 2 3 S  O  •8  o  O v 00 •/-> vo C N >n o cn cn  — "! 00  Tf  m r»  Io O  ov O v  Tf  Tf  U  o o  > • Ocfa  T3 £2 75 00  'SB  O v cn cn  o  Tf tTf Tf  m oo oo cn m r- CN t~ vd 00 od V D vd rr~  Tf  , , vo vo  —-< V O 00  t  ,  »n V D vd >n i n o O o o o C NC NC NC NC N  o H  Appendix B - Selected Shake Flask Data  Summary of Shake Flask Bioleach Experiment Expt. Name: C52 Expt Type: Bioleach (35°C) Substrate: Concentrate CI- Level: 3 gpl Fe(2+) Level: ~9 gpl Revision Date: April 15/95 INPUTS: Solids Ore  Weight (g) Wt% Cu . 2.04 17.30  Wt% Fe 11.60  Wt% S-2 11.62  WtCu  WtFe  WtS-2  (g)  (g)  (g)  0.353  0.237  0.009 0.002 NA NA  0.019 0.676 NA NA  0.364  0.932  0.237  Solutions Inoculum Media Acid 211 gplNaCl  Volume(mL) 5 70 0.55 1.726  System Volume (neglecting acid):  gpl Cu 1.71 0.030 NA NA  gplFe 3.88 9.65 NA NA  76.726  OUTPUTS: Solids Leach Residue  Weight (g) Wt% Cu 3.0500 0.483  Wt% Fe Wt% (S-2+S°) 17.400 2.37  0.237  WtCu  WtFe  Wt% (S-2+S°)  (g)  (g)  (g)  0.015  0.531  0.0019 0.0039 0.0042 0.0043 0.0043 0.3173 0.0216  0.00885 0.00675 0.00480 0.00473 0.00468 0.34725 0.01968  0.372  0.927  Input - Output: -0.008542 % Lost of input: -2.35%  0.0041 0.44%  0.072  Solutions Volume(mL)  gpl Cu  gpl Fe  Sample # 1 2 3 4 5 Filtrate Wash  75 300  1.85 3.88 4.2 4.3 4.3 4.23 0.0720  8.85 6.75 4.8 4.73 4.68 4.63 0.0656 Total OUT (g):  Acid Addition Summary Volume 6M H2S04 added (ml): Acid Consumption (kg H2S04/tonne cone):  0.55 158.66  Copper Extraction Summary Copper Extraction based on Outputs (%): Calculated Head Copper (%):  95.9 17.7  Sulphate Formation Summary Sulphate Formation (%):  69.5  119  0.072  1.726  0.00  0.00  0.00  0.00  o  0.00  u «  1.726  — *o  inoo©©© in V O CN Tf ©©©© ©' X inmTf cnV Ocn Tf Tf T f c/i V O bincinn oo cn© O N in O Oc n TV fC m inin VO VO X C inN Tf CN NV OC © ©cnin ©' © © iinn © T f c n T f C N CNCN TI-f; ca  CN  Tt  o  o  o  m  00  C  00  00  od vd  UH  o 13  vo  oo  vo  oo  vo  o.  00  o o  vj  f- j  C  ^  o  H  o  CD „—,  » .5 Q H  c o  ca  Q  cn  o  § N  o U  fl<  <  U  2 .J E  CO  oo S3  N tCN > in 2 U  a o  i—i  w -J a,  2 <  N  a  w  OV O CN  < Vi  Ov ov  3  T3  o  —1 UJ  < Vi  o  —  3  —<  00  ^  O  g  c oo o u  /-v  -a  >  3fled i> S _e £ >v 3 Vi Vi  i s  cj X  •a  ca  C o U  as =3 ±3  O <U  3 3 S g o  -a  <u  JS  3 o "3 U  >, J: j<  .•3  00  Ov  0 0 0 0 00 vd od  oo vo  ov  5 ..  * *  — o  o  Ov Ov  fe X  e o  ca  vd Ov Ov od od  00  — cn cn P ^  ca  00  Ov Ov Ov  s  u  UH  vo  c  T"j Tf  M  -v° 0 -  K. "  <c  in  oo oo vq n oi  Vi  X  v=  c o S o  Ov  u 3 u, o  T 3 O  cd  Ov Ov  oo  X  c/3  ca  oo  vd  •a c  3  cd  Ov vd vd  UJ  3  C  a. c  3  -a  CQ  CNcn CNCN CNcTnf CN CN ©' ccn V cCn N© CN nO ©'r~O V cm nm © inncn c n c n c CN CNCN ccn N niTnf C Tf cn mcn©cn cn Tf Tf Tf Tf VO c n o o in ©cncncn f Tf Tf"Tf cnTCN CN cnV c n c n c n n O cc~ n ©©© ©cr~ Tf V O Tf V OTf Tf Tf cd  C  T3  r~  Ov Ov So  c  CH  c  m  r~  Tf Tf O TV f Tf Tf Tf V OSoovCN C N T t " O VO V  6  0 2 Vi T3  m m  6  oo o oo  vo vo  vd  00  w  •a > 5=  %  00  r~ vd  vo t~  r~ r~ vd vd vd vd r~ r~ r~  o  od vo  t~  vo vo vo  ©  CN  Appendix B - Selected Shake Flask Data  Summary of Shake Flask Bioleach Experiment Expt. Name: Expt Type: Substrate CI- Level: Fe(2+) Level: Revision Date:  C52ST Sterile (35°C) Concentrate 3 gpl ~9 gpl April 15/95  INPUTS: Solids Ore  Weight (g) Wt% Cu 2.03 17.30  Wt% Fe 11.60  Wt% S-2 11.62  WtCu  WtFe  Wt S-2  (g)  (g)  (g)  0.351  0.235  0.000 0.002 NA NA  0.000 0.676 NA NA  0.353  0.911  0.236  Solutions Inoculum Media Acid 211 gplNaCl  Volume(mL) 5 70 0.7 1.726  System Volume (neglecting acid):  gpl Cu 0 0.03 NA NA  gpl Fe 0 9.65 NA NA Total IN (g):  76.726  OUTPUTS: Solids Leach Residue  Weight (g) Wt% Cu 1.7300 6.030  Wt% Fe Wt% (S-2+S°) 13.400 14.8  0.236  WtCu  WtFe  Wt% (S-2+S°)  (g)  (g)  (g)  0.104  0.232  0.0018 0.0023 0.0025 0.0027 0.0029 0.2318 0.0123  0.00868 0.00845 0.00830 0.00830 0.00825 0.59660 0.03090  0.361  0.901  Input - Output: -0.007329 -2.07% % Lost of input:  0.00968 1.06%  0.256  Solutions Volume(mL) Sample # 1 2 3 4 5 Filtrate Wash  1 1 1 1 1 76 300  gpl Cu 1.8 2.3 2.53 2.68 2.89 3.05 0.0410  gplFe 8.68 8.45 8.3 8.3 8.25 7.85 0.103 Total OUT (g):  Acid Addition Summary Volume 6M H2S04 added (ml): Acid Consumption (kg H2S04/tonne cone): Copper Extraction Summary Copper Extraction based on Outputs (%): Calculated Head Copper (%):  0.70 202.92  70.9 17.7  Sulphate Formation Summary Sulphate Formation (%):  121  0.256  -  « 2 1: O  vo O  (N o o  o o o  o o o  o o o  o o o  m o  o  o  o  o o  m o o o  H - "7 13  CN  VO  oo g  g  CN  H H ©  H H  o  o o  53 oo in V D 0 0 00  a. u.  rn m m in 00 od CN oo od  3 o  ^ & oo 6 0  OO rn m oo O V in O OO o CN in V CN CN CN rn 00  oo o V O 00 VO VO v ov in 00 O in in in m in m  tn  in CN VO VO Vo ov O CN  o o o  CN V O CN V OV O H — CN o O cn in vd r~ o  o  in in o in o CN CN o in r~ rn CN m  53  o  g  .5 & 3  H _j H  8 CS  Q  o n e  c n ca  E  •=r  CS  _  O vO vO VO VO V gv oo m V O OO H H CN CN CN CN CN >> >> >, >s >> C3 ca ca ca ca ca  J3  00  "5  r-n  oo  - cu  i« CQ X  -o c o  s ON i  o 00 CN vq t~ in rn in CN  O  O.  T3  erf  O.  ° w  as O  •a ii 5 n ca D ~  CM 2 O N t—  E  o g  w 33  O v •sj- cn vq O v t~ O vd od rn od m in in in vo vo  <D  00  t~ C2 •  in vo vo ov —  O  o  J uj —i  3  2  E  <  o ~". <-> • * CN  <  3  O  —  u  r-  w  . .  ~ 1 00 w  "S3 ^ C  m  S  3  » (X,  o tc o So  1' = ~  O v >v O  5  t-; 00 H ^  >n m  CN m in O V OV O in V  oo O v  X  -5  OH  ac o <c -j o  —  T3  ca g  O vO vO v in V vd O vO CN V OV O vo  M  ^  2 E  4) o  1=1  i  121  HJ  u_  > g  § 2  i  >,•§§! oo oo £  2  o  x  .5  o  S3 "9  S > > .  H C  3 ' = & B  CO  U  O  v vo o 0 0 O 0 0 rn vq vq 0 0  m o CN CN CN CN rn  m  CN  w  O  T3  8  00 V O o m in V D vq in rn CN oq 00 od od od od r~  m  m  m  vq O v i> r~ vo 0 0 0 0 vd vd vd  r - r~ r~ r» o  V oo oo oo oo r - O H H o m H H H H H H vo vd  o o o o o o CN CN CN CN CN CN  H  Appendix C - Selected Batch Reactor Data  Summary of Batch Reactor Bioleach Experiment Expt. Name: Expt Type: Ore: CI- Level: Fe(2+) Level: Revision Date:  BO 11 Bioleach - pulp density 10% Ore 3 gpl ~9 gpl April 15/95  INPUTS: Solids Ore  Weight (g) Wt% Cu 199.40 1.56  Wt% Fe 2.6<  Wt% S-2 0.81  WtCu  WtFe  WtS-2  (g)  (g)  (g)  3.111  5.304  0.235 0.000 NA NA  0.598 16.489 NA NA  3.346  22.391  1.615  Solutions Inoculum Media Acid 211 gplNaCl  Volume(mL) 130 1820 3.9 44  gpl Cu 1.81 0 NA NA  gpl Fe 4.6 9.06 NA NA  1994  System Volume (neglecting acid):  Total IN (g):  OUTPUTS: Solids Leach Residue  Weight (g) Wt% Cu 236.4900 0.150  Wt% Fe Wt% (S-2+S°) 6.120 0.24  1.615  WtCu  WtFe  Wt% (S-2+S°)  (g)  (g)  (g)  0.355  14.473  0.0008 0.0011 0.0013 0.0014 0.0014 2.7383 0.0551  0.00894 0.00863 0.00565 0.00465 0.00428 8.02775 0.23617  Total OUT (g):  3.154  22.769  Input - Output: % Lost of input:  0.192 5.73%  -0.378 -1.69%  0.568  Solutions Volume(mL)  gpl Cu  gpl Fe  Sample # 1 2 3 4 5 Filtrate Wash  Acid Addition Summary Volume 6M H2S04 added (ml): Acid Consumption (kg H2S04/tonne cone):  1970 2090  0.835 1.14 1.32 1.36 1.37 1.39 0.0264  8.94 8.63 5.65 4.65 4.28 4.075 0.113  3.90 11.51  Copper Extraction Summary Copper Extraction based on Outputs (%): Calculated Head Copper (%):  87.8 1.46  Sulphate Formation Summary Sulphate Formation (%):  64.9  123  0.568  o  — «  o o  —Z  22  S  "8 -  CN  5  VO  o  o  o  o  o  o o o  o o o  o o o  o o o  o o o  in  00  o o  o o  o o  o o  o o  o o o  "a-  ^  o  OS  ro  -O O  CN  ac *  •* r o Ov  in  vq  vq  VO  00 o d  in  •*}•'  CN  in  r-  o  3 o  H in .£  ro  3  •a- C N V O ro  00  OH ( J  CO  ro  OV CO  CO  o  o  VO  o  Ov  «-)V O  00 00 o o  oo oo  _ Ov >—i  Ov oo  in  oo  vo  OV  OV  OV  vo  CN  vo  VO  in  in  CN  O  o o  43  H  ca u  ed  o  D  0 ca ca  CQ  in  ro  in in  in  vd  .5  in  CN  O in  o in  o o  o o  ro  CO  •a- •a-  CO CN  CN  CN  •a-  -a-  Q H  l-  <D 01  00 v q  —1 o  a  O  l a .  £> ov  OV  H H CN  CN  00 o  6 2  s  60 3  60 3  cs  60 3  Ov  OV  CO CN  CN  60 3  "aOv  in  60 3  CN  60 3  << << < <  is » 1  o  u '"S e ca  o a. a.  so  -a c o  3<3  O -a  CQ N  Os  O0 C N O OV CN i n 00 o o  VO  Ov  CO  Ov  vd  o  o  in  O  - S ° _" >  X 2 O 13  gj  5 H O <-> SC  -a c  2  U-  ^  0 -  VO  VO 0  1 o  O Ov Os  O  O  *3* c o e*. —  00  r~ t—  3  r°o  VO  vd i n  CN  5  CJ  H  in  r~  „  CO  od CO  vd CO  oo  u  •a-  VO  r~  -aOs  ro vq  vq  od  od  in  CN  J3  in  oo  <*  . ^  Ov  c  s  00 o  S  3  o  •a- v o  o  m  Os  o  HJ  t~in  VO  oo  oo  00 o o CN  o  •a-' •a-'  fa  £•  •a- •a- C N V O o o —* C O C O 3  U  O  »  ob  CO  Ov CO  o  o  o  o  5 0.  o H U  <  a a  u  < CQ  o  + cr O  3  JH  31  "o 5  >  — — .s?  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