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Kinetics of leaching of chalcocite in acid ferric sulfate media : chemical and bacterial leaching Bolorunduro, Samuel Adewale 1999

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KINETICS OF LEACHING OF CHALCOCITE IN ACID FERRIC SULFATE MEDIA: CHEMICAL AND BACTERIAL LEACHING by SAMUEL ADEWALE BOLORUNDURO B.Sc.(Hons.), Obafemi A w o l o w o University, Ile-Ife, 1990  A THESIS S U B M I T T E D IN P A R T I A L F U L F I L L M E N T OF T H E R E Q U I R E M E N T S F O R T H E D E G R E E OF M A S T E R OF A P P L I E D S C I E N C E in T H E F A C U L T Y OF G R A D U A T E S T U D I E S Department o f Metals and Materials Engineering  We accept this thesis as conforming to the required standard /I  T H E U N I V E R S I T Y OF BRITISH C O L U M B I A September, 1999 ©Samuel Adewale Bolorunduro  In  presenting  degree freely  at  this  the  University  available  copying  of  department publication  of  in of  reference  for  this or  thesis  thesis by  this  for  his  fulfilment  British  Columbia,  and  study.  scholarly her  or  thesis  partial  for  I  of I  further  purposes  gain  shall  requirements  agree  that  agree  may  representatives.  financial  the  It not  be  the  that  Library  by  be  allowed  make  head  that without  of  my  copying  or  my  permission.  Department  of  T h e U n i v e r s i t y of British Vancouver, Canada  DE-6  (2/88)  Ku*  k Hkltt-JK^S  Columbia  fc*Tft  I ^  f  c  1  it  extensive  for  the  understood  advanced  shall  permission  granted  is  an  for  ^9  written  ABSTRACT The lack o f a clear understanding o f the rate o f chalcocite ( C u S ) and covellite 2  (CuS)  leaching in the presence and absence o f bacteria has been a limitation on the  optimization o f hydrometallurgical processes for the recovery o f copper from these minerals. In order to enhance the performance o f heaps and other leaching processes for these minerals, there is a need to examine the conditions required to improve the rate o f leaching. Such an investigation w i l l produce a particle scale model (for the intrinsic rate of leaching), which can be combined with a heap scale model to form a comprehensive heap leaching model, or with a leaching macro-model to form a tank leaching model. The robustness o f such a model w i l l be its ability to determine the effects o f changing parameters on the rate limiting steps. Chalcocite oxidation was investigated by leaching high grade natural minerals in acidic ferric/ferrous sulfate solutions. The temperature was varied between 35° and 75° C . The ferric concentration varied between 0.015 and 0.232 m o l / L , and the  ferrous  concentration varied between 0.001 and 0.233 mol/L. The redox potential o f the solution (at 25°C) was varied between 450 and 651 m V (vs. A g / A g C l ) to determine the effect o f this parameter on the leaching kinetics. The two well  known, significant stages o f leaching were  observed  and  characterized by mineralogical studies. The first stage leach was characterized by 50% copper extraction and the conversion o f chalcocite, ultimately to second stage covellite (CuS). Some non-stoichiometric copper sulfides were formed prior to the formation o f the second stage covellite. The first stage leaching reaction was rapid at all temperatures. The redox potential had no effect on the rate o f this reaction. The second stage leach was characterized by the conversion o f the second stage covellite (which was the by-product o f the first stage) to copper, elemental sulfur and sulfate. A t higher temperatures, sulfur formation was predominant and the reaction was fast. A t lower temperatures, sulfur formation was predominant up to about 70% copper extraction. Subsequently, sulfate formation occurred. The effect o f the solution redox  u  potential on the kinetics o f second stage leaching was significant. During bacterial leaching, it was observed that the principal role o f the ferrous oxidizing  bacteria  (Thiobacillus ferrooxidans) was to maintain the required high redox potential at the surface o f the minerals. A mathematical model was formulated to explain the ferric leaching kinetics o f chalcocite.  The first  stage  kinetics  can be  explained  i n terms  of a  mixed  diffusion/chemical reaction model, i n which the rate o f reaction is simultaneously limited by the diffusion o f ferric ions to the mineral surface and by the chemical reaction. Though the partially oxidized particles disintegrate before the commencement o f the second stage leach, each o f the particles leaches as a discrete grain and the second stage kinetics are controlled by the chemical reaction, which is one-half order dependent on the ferric concentration. The leaching process can be described by an electrochemical mechanism in which the rate-limiting step o f the first stage is electron transfer i n the cathodic reaction, and the rate-limiting step o f the second stage is electron transfer i n the anodic reaction.  111  TABLE OF CONTENTS ABSTRACT  ii  T A B L E OF CONTENTS  iv  LIST OF T A B L E S  vi  LIST OF FIGURES  vii  LIST OF S Y M B O L S  viii  C H A P T E R 1: I N T R O D U C T I O N  1  C H A P T E R 2: L I T E R A T U R E R E V I E W  4  2.1 Properties of Chalcocite and Covellite 2.1.1 Crystal Structures 2.1.2 Leaching Implications of the Structures 2.1.3 Thermodynamics Considerations 2.1.4 Multiple Stage Dissolution o f Chalcocite 2.2 Polarization Behaviour of Chalcocite and Covellite 2.2.1 Effect o f Electrolytes on Polarization 2.2.2 Effect of Temperature on Polarization 2.3 M i x e d Potential Theory of Leaching 2.3.1 Type I Leaching 2.3.2 Type II Leaching 2.3.3 Type III Leaching 2.4 Application of M i x e d Potential Theory to Leaching 2.4.1 Leaching o f Chalcocite 2.4.2 Leaching o f Covellite 2.5 Chemical Leaching Kinetics 2.6 Effect o f Parameters on the Kinetics in Sulfate M e d i a 2.6.1 Effect o f Stirring Speed 2.6.2 Effect o f Temperature 2.6.3 Effect o f Particle Size 2.6.4 Effect o f Ferric Concentration 2.6.5 Effect o f Acidity 2.7 Effect o f Parameters on the Kinetics in Chloride M e d i a 2.7.1 Effect of Temperature 2.7.2 Effect of Ferric Concentration 2.7.3 Effect o f Particle Size 2.8 Proposed Mechanisms of Sulfate Leaching 2.8.1 Cathodically-Controlled M i x e d Potential Mechanism 2.8.2 Iron Depassivation M i x e d Potential Mechanism 2.9 Proposed Mechanisms of Chloride Leaching  iv  4 4 7 7 11 13 17 21 22 26 27 29 30 30 31 34 37 37 38 38 39 39 40 40 41 41 41 42 44 47  2.10 Chalcocite Leaching in Other M e d i a  53  2.11 Summary of Literature Review  53  C H A P T E R 3: E X P E R I M E N T A L P R O C E D U R E S  55  3.1 Chemical Leaching Experiments  55  3.1.1 Sample Preparation and Minerals 3.1.2 Chemical Analysis of the Sample 3.1.3 Mineralogical Characterization of the Sample 3.1.4 Selection of Monosize Particles 3.1.5 Chemical Leaching Apparatus and Procedures 3.1.6 Role o f Potassium Permanganate 3.1.7 Slurry Filtration Method 3.1.8 Analytical Methods 3.1.9 Determination of Copper Extractions and Sulfide Oxidation 3.2 Mineralogical Characterization of Reaction Products 3.2.1 Qualitative Analysis 3.2.2 Phase Compositional Analysis 3.3 Bacteria leaching of Chalcocite 3.3.1 Bacterial Culture and Nutrient M e d i a 3.3.2 Bacterial Leaching Experiment  55 57 57 58 59 61 65 65 66 67 67 67 68 68 69  C H A P T E R 4: R E S U L T S A N D D I S C U S S I O N 4.1 Mineralogical Characterization 4.1.1 X-ray and Microscopic Analyses of Feeds 4.1.2 X-ray Analyses of Leached Residues 4.1.3 Qualitative Analyses o f Leached Residues by S E M - E D X 4.1.4 Compositional Changes by Electron Microprobe Analysis 4.2 Results o f Chemical Leaching Experiments: First Stage Leaching 4.2.1 Effect of Controlled Potential 4.2.2 Effect of Temperature 4.2.3 Effect of Initial Ferric Concentration 4.2.4 Effect o f Particle size 4.2.5 Effect of Initial Ferrous Concentration 4.3 Results o f Chemical Leaching Experiments: Second Stage Leaching 4.3.1 Effect of Controlled Potential 4.3.2 Effect of Temperature 4.3.3 Effect of initial Ferric Concentration 4.3.4 Effect of Initial Particle size 4.3.5 Effect of Initial Ferrous Concentration 4.3.6 Effect of Ferric/Ferrous Ratio 4.4 Effect of Leaching Parameters on Sulfur Distribution 4.5 Bacterial Leach Experiment C H A P T E R 5: T H E O R Y A N D M O D E L I N G  71 71 72 76 78 85 88 88 89 90 92 95 98 98 99 101 102 103 107 110 113 117  5.1 Physico-Chemical M o d e l of First Stage Leaching  v  117  5.2 5.3 5.4 5.5  Physico-Chemical Model o f Second Stage Leaching Electrochemical M o d e l o f Second Stage Leaching Rate Expression for First Stage Leaching Rate Expression for Second Stage Leaching  120 121 123 126  C H A P T E R 6: C O N C L U S I O N S  133  C H A P T E R 7: R E C O M M E N D A T I O N S F O R F U R T H E R W O R K  135  REFERENCES  136  A P P E N D I X 1: E H - P H D I A G R A M A T H I G H T E M P E R A T U R E  143  A P P E N D I X 2: E X P E R I M E N T A L D A T A  147  A P P E N D I X 3: P O T A S S I U M P E R M A N G A N A T E A D D I T I O N  153  A P P E N D I X 4: T O P O C H E M I C A L M E C H A N I S M O F L E A C H I N G  157  vi  LIST OF T A B L E S Table 2-1. Kinetics o f first stage ferric leaching o f C u S  35  2  Table 2-2. Kinetics o f second stage ferric leaching o f C u S and CuS 2  36  Table 3-1. Results o f chemical analyses for the as-received and reground minerals ..57 Table 3-2. Results o f mineralogical composition o f the chalcocite sample  58  Table 4-1. Summary o f qualitative analyses o f the leached residues by S E M .  82  Table 4-2. Chemical compositions o f the resulting phases by E P M A  86  Table 4-3. Chemical compositions o f the minor phases by E P M A  87  Table 4-4. Geometric analysis o f the different Particle size fractions.  92  Table 4-5. Effect o f redox potential on sulfur distribution i n the leach residue  110  Table 4-6. Effect o f temperature on sulfur distribution in the leach residue  112  vii  LIST OF FIGURES Figure 2-1.  Crystal structure, bond lengths and bond angles of chalcocite  6  Figure 2-2.  Crystal structure of covellite  6  Figure 2-3.  E h - p H diagram for the C u - S - H 0 System at 25 °C, activity o f copper 2  ion at 0.01 mol/L and other ionic species at unit activity Figure 2-4.  8  E h - p H diagram for the C u - S - H 0 System at 75°C, activity of copper 2  ion at 0.01 mol/L and other ionic species at unit activity  8  Figure 2-5.  Anodic polarization curves for synthetic C u S and CuS electrodes  14  Figure 2-6.  Anodic polarization behavior o f a pre-oxidized C u S electrodes  15  Figure 2-7.  Anodic polarization behavior o f a C u S and pre-oxidized C u S electrodes  15  Figure 2-8.  Passivation behaviour of massive and particulate C u S anodes  16  Figure 2-9.  Passivation behaviour o f chalcocite anodes i n cupric sulfate and cupric chloride electrolytes  18  2  2  2  Figure 2-10. Passivation behaviour o f chalcocite anodes in H S 0 and N a C l 2  4  electrolytes  19  Figure 2-11. Effect o f chloride additions on the passivation behaviour o f C u S 2  anodes i n a sulfate electrolyte  20  Figure 2-12. Effect o f temperature on the passivation behaviour o f particulate C u S 2  anodes  22  Figure 2-13. Current density potential diagram showing curves for metal sulfide with rest potential E , and oxidant with rest potential E e  e 2  25  Figure 2-14. Polarization curve for Type I leaching  27  Figure 2-15. Polarization curve for Type II leaching  28  Figure 2-16. Polarization curve for Type III leaching  29  viii  Figure 2-17. Polarization curves for anodic dissolution o f C u S (0.1 m o l / L C u S 0 , 2  4  0.5 m o l / L M g S 0 , 1 m o l / L H S 0 ) and cathodic reduction o f ferric 4  2  4  ions on graphite (0.1 mol/L Fe2+, 1 mol/L H S 0 4 ) at 25°C  31  2  Figure 2-18. Polarization curves for anodic dissolution o f CuS (0.1 mol/L C u S 0  4  and 0.1 mol/L H S 0 ) and cathodic reduction of ferric ions on graphite 2  4  at 55°C  33  Figure 2-19. Effect of temperature on the leaching of chalcocite ( F e C l , 0.5 m o l / L , 3  HC1 0.2 mol/L, F e / F e 3 +  2 +  ratio 10, 2.5 g of+150-300 urn C u S ) 2  40  Figure 2-20. Illustration of half-cell and mixed potential variations during ferric leaching of chalcocite: 0.1 mol/L C u S , 0.5 mol/L F e 2  0.001 mol/L F e  2 +  and C u  3 +  and 44  2 +  Figure 2-21. Schematic Evans diagram o f C u S pressure leaching 2  46  Figure 2-22. Effect o f chloride concentration on dissolution, at 303 K , 0.086 M P a oxygen pressure, 0.35 mol/L [H ] and 210 x 177 urn +  48  Figure 2-23. Effect o f temperature on the second stage leaching at 0.5 m o l / L C f 0.086 M P a pressure, 0.35 mol/L [ H ] and 210 x 177 jam particles +  49  Figure 2-24. Physico-chemical model for first stage chalcocite dissolution in the sulfate system  50  Figure 2-25. Physico-chemical model for first stage chalcocite dissolution i n the chloride system  51  Figure 2-26. Physico-chemical model for second stage chalcocite dissolution in the sulfate system  52  Figure 2-27. Chalcocite dissolution in various media Figure 3-1.  54  Cumulative particle size distributions o f the mineral samples as determined by screening with Tyler sieves  56  Figure 3-2.  Schematic representation of the controlled potential Leaching set-up  61  Figure 4-1.  X-ray diffraction pattern of chalcocite powder  72  ix  Figure 4-2.  Figure 4-3.  Figure 4-4.  X-ray display (elemental composition) analysis o f a chalcocite rock section  73  Backscattered electron image o f a rock section o f covellite showing chalcocite zoning in covellite  74  X-ray display (elemental composition) analysis o f a covellite rock section  74  Figure 4-5.  X-ray diffraction pattern o f natural covellite powder  75  Figure 4-6.  Backscattered electron image o f -250+212 p m grains o f chalcocite showing chalcocite (1), pyrite (2) and bornite (3) at 0% extraction  76  Figure 4-7.  X-ray diffraction pattern o f the leached residue after 44% extraction from -250+212 p,m grains o f chalcocite at 35°C, 0.116 m o l / L ferric  Figure 4-8.  Figure 4-9.  and 0.020 mol/L ferrous concentration  77  Backscattered electron image o f the leached grains (of chalcocite) after 30% extraction at 35°C, showing cracks, pores and breakage o f particles. The identified grains are copper-bismuth-sulfide (1) and covellite (2)  79  Digital image o f a leached grain showing cracks and pores after 44% extraction  79  Figure 4-10. Backscattered electron image o f a leached grain o f chalcocite showing puffy and fragile textures after 30% extraction ....80 Figure 4-11. Backscattered electron image o f selected leached grains for the microprobe (compositional) and elemental analysis o f phases, after 44% extraction  80  Figure 4-12. Backscattered electron image o f selected leached grains after 10% extraction  81  Figure 4-13 Backscattered electron image o f selected leached grains after 40% extraction  81  Figure 4-14. Backscattered electron image o f a yarrowite grain showing cracks  83  Figure 4-15. X-ray analysis o f a C u - B i - S phase, showing some paragenics effects  83  x  Figure 4-16. Backscattered electron image o f an iron zone (1) i n copper-sulfide phase (2)  84  Figure 4-17. Backscattered electron image o f adjacent grains o f bornite and pyrite, showing the preferential leaching o f the bornite grain  85  Figure 4-18. Effect o f potential on the first stage leaching o f - 250+212 u m particles at 35°C0.116 mol/L ferric and 0.0202 mol/L ferrous  88  Figure 4-19. Effect o f temperature on the first stage leaching o f -250+212 um particles at 0.116 mol/L ferric and 0.0202 mol/L ferrous  89  Figure 4-20. Arrhenius plot for the first stage leaching  90  Figure 4-21. Effect o f ferric on the first stage leaching o f -250+212 um particles at 35°C and 0.0101 mol/L ferrous  91  Figure 4-22. L o g rate vs. log ferric concentration plot for the first stage leaching  92  Figure 4-23. Effect o f particle size on the first stage leaching at 35°C, 0.116 m o l / L ferric and 0.0202 mol/L ferrous Figure 4-24. Rate vs. the reciprocal o f radius squared plot for first stage leaching  93 .....94  Figure 4-25. Rate vs. the reciprocal o f radius plot for first stage leaching  94  Figure 4-26. Effect o f ferrous on the first stage leaching o f -250+212 um particles at 35°C and 0.116 mol/L ferric  95  Figure 4-27. Effect o f ferrous on the first stage leaching o f -250+212 jam particles at 35°C and 0.232 mol/L ferric  96  Figure 4-28. L o g rate vs. log ferrous concentration plot for the first stage leaching of-250+212 u.m particles at 35°C and 0.232 mol/L ferric Figure 4-29. Reaction order vs. ferrous concentration plot for the first stage leaching  97  97  Figure 4-30. Effect o f potential on the second stage leaching o f -250+212 (am particles at 75°C, 0.116 mol/L ferric and 0.0202 m o l / L ferrous  98  Figure 4-31. Effect o f temperature on the second stage leaching o f -250+212 jam particles at 75°C, 0.116 mol/L ferric and 0.0202 mol/L ferrous  xi  99  Figure 4-32. Arrhenius plot for the second stage leaching  100  Figure 4-33. Effect o f initial ferric on the second stage leaching o f -250+212 p m particles at 75°C and 0.0101 mol/L ferrous  101  Figure 4-34. L o g rate vs. log ferric concentration plot for the second stage leaching,.... 102 Figure 4-35. Effect o f particle size on the second stage leaching at 75°C, 0.116 m o l / L ferric and 0.0202 mol/L ferrous  103  Figure 4-36. Effect o f ferrous on the second stage leaching o f -250+212 p m particles at 75°C and 0.116 mol/L ferric  104  Figure 4-37. L o g rate vs. log ferrous concentration plot for the second stage leaching o f -250+212 p m particles at 75°C and 0.116 m o l / L ferric  105  Figure 4-38. Reaction order vs. ferrous concentration plot for the second stage leaching at 75°C  106  Figure 4-39. Effect o f ferrous on the second stage leaching o f -250+212 p m particles at 35°C and 0.116 mol/L ferric  106  Figure 4-40. L o g rate vs. log ferrous concentration plot for the second stage leaching of-250+212 p m particles at 75°C and 0.116 m o l / L ferric  107  Figure 4-41. Effect o f ferric/ferrous ratio on the second stage leaching o f -250+212 p m particles at 75°C, 0.136 mol/L total iron and 0.287 m o l / L sulfate  108  Figure 4-42. L o g rate vs. log o f ferric/ferrous ratio for the second stage leaching at 75°C  109  Figure 4-43. Reaction order vs. ferric/ferrous ratio plot for the second stage leaching  109  Figure 4-44. Effect o f bacteria on the leaching (first and second stage) of-250+212 p m particles at 35°C and 0.116 mol/L total-iron. The initial concentration o f ferric was 0.0172 mol/L  115  Figure 4-45. Rate vs. extraction plot for the bacterial and chemical leaching (second stage) o f -250+212 p m particles at 35°C and 0.116 m o l / L total iron  xii  116  Figure 4-46. Redox potential and p H profile o f bacterial leaching and chemical leaching  116  Figure 5-1. Physico-Chemical model for the first step o f first stage leaching  117  Figure 5-2. Physico-Chemical model for the second step o f first stage leaching  118  Figure 5-3. Relationship between rate and cuprous concentration for the first stage leaching o f -250+212 p m particles at 35°C, 0.116 m o l / L ferric and 0.0101 mol/L ferrous ..:  119  Figure 5-4. Physico-Chemical model for the second stage leaching  120  Figure 5-5. Evans diagram o f theoretical polarization curves during ferric sulfate leaching o f second stage covellite  122  Figure 5-6. L o g rate vs. log o f extraction for the second stage leaching o f 250+212 p m particles at 65 and 75°C, 0.116 mol/L ferric and 0.0202 m o l / L ferrous  129  Figure 5-7. Predicted vs. actual extraction for the second stage leaching o f 250+212 p m particles at 75°C and different ferric / ferrous ratio  131  Figure 5-8. L o g rate vs. log o f extraction for the second stage leaching o f 250+212 p m particles at 55°C, 0.116 mol/L ferric and 0.0202 mol/L ferrous  132  Figure 5-9. Predicted vs. actual extraction for the second stage leaching o f 250+212 p m particles at 55°C, 0.116 mol/L ferric and 0.0202 m o l / L ferrous  132  xiii  LIST OF S Y M B O L S G°  Gibbs free energy at a temperature (T), kJ/mol  T  temperature, ( K )  C  heat capacity, J/mol-K  a, b,c  heat capacity coefficients for the non-ionic species  a •, fi 7  T  Criss-Cobble heat capacity constants for the ionic species  S  entropy, J/mol-K  S  absolute entropy, J/mol-K  E  reversible (or equilibrium) potential o f the half cell, V  e  EoxE  mixed-potential o f the system, V  V  overpotential  i orj  current density  F  faraday, 96485 C/mol e"  fi  electron (or charge) transfer coefficient  k  rate for the forward direction  k  rate for the reverse direction  D  diffusivity, cm /min  5  boundary layer thickness, cm  C11  calculated head assay o f copper, g.%  M  ^"calc  mass o f copper solubilized, g,  <;  mass o f copper in the leach residue, g mass o f the head sample, g  '"sulfide  mass o f sulfide sulfur in the head sample, g  res "'sulfide  mass o f sulfide in the leach residue, g  res '"sulfur  mass o f elemental sulfur i n the leach residue, g  ^*90  90% passing through  xiv  ACKNOWLEDGEMENTS I would like to express my sincere gratitude to my supervisors, D r . Ralph H a c k l and Dr. D a v i d D i x o n , for their constant guidance and support throughout the course o f my study. It was the excellent initiative o f Dr. Ralph H a c k l which made it possible for me to pursue my masters degree in the University o f British Columbia. Thanks are also extended to my wife (my pillar o f inestimable strengths), whose understanding and encouragement during the trying times were appreciated. I would like to extend my appreciation to my fellow graduate students and the research engineers i n the Hydrometallurgy Group, especially to Dr. Berend Wassink for his assistance during the experimental setup. The financial assistance from the C y and Emerald Keyes Foundation and the University o f British Columbia Graduate A w a r d is gratefully acknowledged.  xv  CHAPTER 1 INTRODUCTION The mineral chalcocite is the most abundant copper sulfide after chalcopyrite, and the most amenable to hydrometallurgical treatment. Bacterial heap leaching, both with and without prior agglomeration and acid curing, is one o f the most important emerging technologies for the treatment o f l o w grade secondary copper sulfide, and mixed sulfide/oxide o f copper ores. The process is popular especially i n South America, because o f its simplicity and low cost o f operation. Regrettably, the process has been hindered by very l o w rates o f extraction and lower ultimate copper extractions than expected. This poor performance is attributed to the slow kinetics o f the second stage o f chalcocite oxidation. The problem is compounded by the lack o f clear understanding o f the kinetics o f chalcocite leaching. The various non-stoichiometric copper sulfide compounds, w h i c h are formed as a result o f successive removal o f copper atoms from the initial chalcocite matrix, have been blamed for the slow kinetics o f this stage. It is also believed that the leach path (i.e., the transformation to different non-stoichiometric copper  sulfides)  changes with solution chemistry (such as ferric ion concentration). The leaching o f chalcocite has been investigated under different conditions i n the laboratory for seventy years. The investigations began with Sullivan [1] i n 1930 and several researchers [2-13] have since studied chalcocite leaching by using ferric ion as the oxidant i n either a sulfate or chloride environment. M a n y o f the laboratory studies have used electrochemical oxidation to shed light upon the phase transformations which occur when copper is removed from chalcocite and to explore the anodic dissolution o f chalcocite as a recovery method for copper [14-17]. A l s o , some studies have investigated the oxidation o f chalcocite by Thiobacillus ferrooxidans bacteria i n oxygenated and nonoxygenated acid solutions [18-35]. It is widely known that chalcocite oxidation occurs roughly i n two distinct stages. First stage:  Cu S 2  Second stage: C u S  > CuS + C u >  Cu  2 +  + S°  2 +  + 2 e"  + 2 e"  1  (1-1) (1-2)  The previous kinetic studies o f chalcocite and covellite leaching by ferric [36] have revealed that the first stage leaching o f chalcocite is very fast relative to the second stage leaching. It is generally accepted that the rate o f first stage leaching is controlled by the diffusion o f oxidant to the mineral surface. This view is supported by the l o w activation energies (between 1 and 3 kcal/mol [36]) and the first-order dependence on ferric concentration. O n the other hand, second stage leaching o f chalcocite (or the covellitic phase) is slow and is thought to be controlled by the rate o f the chemical reaction. This view is supported by the high activation energies (between 15 and 25 kcal/mol [36]) and the small dependence o f the rate on ferric concentration at low levels, and very little or no dependence at high levels. These studies were carried out i n the absence o f bacteria, and it is worth noting that no attempts were made to control the redox potential, which would be expected to fall from time zero in every case as oxidant is consumed. This may explain some o f the discrepancies previously reported with regard to the dependence o f rates on ferric concentration and other parameters. Hence the primary objective o f the present work was to control the redox potential i n each experiment, and thereby to establish a clear understanding o f the leaching kinetics o f chalcocite. In addition to quantifying the effects o f ferric ion, temperature, ferrous ion, and particle size, the work also focused on confirming or disproving the hypotheses that the rate o f second stage leaching is governed by the redox potential (as established by the ferric / ferrous ratio). The reasons for the resistance o f the second stage to ferric leaching at l o w temperatures has never been clearly established. The initial rate o f reaction always declines at about forty percent copper extraction, which has led to the conclusion that a reaction product (blue remaining covellite Cuj S ) is formed prior to the formation o f 2  covellite (CuS) and that a long period o f time is required for this transition. Therefore, one o f the objectives o f the present work was to undertake mineralogical characterization o f the reaction products formed prior to second stage. Some investigators believe that a passivating layer o f sulfur or some other oxidation products are formed on the mineral surface during the second stage leaching, and this is corroborated by some electrochemical studies. In the present study, although the focus was not the mineral surface, chemical analyses o f the entire solid product by  2  using an electron-probe microanalyzer ( E P M A ) were helpful i n the investigation o f the passivation phenomenon. The morphological analyses o f the intermediate phases by using a scanning electron microscope ( S E M ) were helpful i n interpreting the leaching curves obtained. The mechanisms by which bacteria such as Thiobacillus ferrooxidans promote the leaching o f chalcocite ( C u S ) and covellite (CuS) have been the subject o f considerable 2  controversy. The two mechanisms which have been proposed are the "direct" mechanism (sulfides oxidized directly by oxygen) and the "indirect" mechanism (sulfides oxidized by bacterially-generated ferric). In any case, the presence o f an active bacterial population is absolutely essential to the success o f copper sulfide heap leaching as currently practised. In this work, a preliminary attempt was made to confirm or disprove the hypothesis that the role o f bacteria is simply to maintain a high redox potential i n the vicinity o f the mineral surface. The approach taken was twofold. First, the entire time history o f the leaching kinetics i n the presence and absence o f bacteria was studied. The kinetics were defined and quantified independently by varying one variable at a time while keeping other ratedetermining parameters constant. The minerals were leached i n batch-wise mode while the solution redox potential was controlled. This concept is termed "controlled potential leaching". The redox potential, feed types, retention time, temperature, p H , ferric and total iron concentrations were monitored independently. Second, the oxidation products were characterized qualitatively by scanning electron microscope and quantitatively by electron-probe microanalysis. Finally, mathematical models were formulated to describe the first and second stage leaching kinetics o f chalcocite.  3  CHAPTER 2 LITERATURE REVIEW 2.1  Properties of Chalcocite and Covellite  The crystal structures o f chalcocite and natural covellite have been inconsistently reported i n the literature. This can be attributed to the similarity between the phases, which are formed during leaching or geological transformation o f the minerals.  2.1.1 Crystal Structures  The crystal structure o f chalcocite, which is illustrated in Figure 2-1 [37] has been widely disputed. This is partly attributed to the closeness o f its structure to that o f djurleite ( C u  197  S ) , which occurs naturally and as an intermediate product during the  leaching o f chalcocite. It can be considered as several minerals and solid solutions whose relations are not well understood  [38]. These relations are complicated by high  temperature transformations, although the low temperature form occurs more frequently than the high temperature form. The three phases that can be identified are the linear copper  co-ordination, triangular  copper  co-ordination  and  tetrahedral copper  co-  ordination [39]. The copper atoms are mainly i n triangular, three-fold co-ordination with a Cu-S bond length o f 2.3 A and S-Cu-S angles ranging from 111.8° to 131.6°. The linear and tetrahedral co-ordinations could result from crystal distortions. The position o f a copper atom could be imposed by another copper atom. The distance between the copper atoms was reported to be about 2.52 A [39], which compared favourably with the C u - C u distance o f 2.556 A i n metallic copper (with a resistivity o f 1.7xl0" Q m ) . Based on this, 8  a metal-metal interaction or stability by metal-metal bonds is suggested [38], though it had been previously concluded that chalcocite is a p-type semi-conductor with resistivity varying from 2 x l 0 "  4  to 1.0 Q m [38]. The structure o f low-temperature  4  chalcocite is  considered as interstitial compounds o f copper i n an approximate hexagonal-closestpacked framework o f sulfur. The reported Cu-S bond length is less than the sum o f the Cu  and S " ionic radii o f 2.80 A , but corresponds to the sum o f the covalent radii. I f  +  2  covalency were ignored, the copper-deficient chalcocite minerals could be represented as Cu _ Cu +  2  2 x  2 +  S " and each missing C u can be considered as an acceptor defect. 2  x  +  In the covellite crystal structure, which is illustrated in Figure 2-2 [40], copper occurs i n both triangular and roughly tetrahedral co-ordination and four o f the sulfur atoms are combined i n S "dianion units. The formula which has been proposed for 2  covellite is ( C u ) ( C u ) 2 ( S ) " 2 ( S " ) 2 in which C u IV  +  III  4  2+  2  2  2  2 +  occupies triangular and C u  +  occupies tetrahedral sites [41]. The S group has a bond length o f 2.071 A indicating a 2  strong covalent bond and the strong bond in the S group requires an average valence o f 2  copper to be less than 2. The average bond length i n the Cu(2+)-S tetrahedron is 2.312 A , which is slightly greater than that o f chalcopyrite and for Cu(l+)-S, it is 2.190 A (Figure 2-2). The thermal motion is markedly varied and anisotropic for the different atoms i n the covellite. Natural covellite is ap-type semiconductor, with a resistivity o f 7x10"  Qm.  The valence o f copper in covellite is still a subject o f dispute, however an X - r a y photoelectron spectroscopic study has shown that copper is monovalent i n most sulfides, including covellite [42]. In another study, the same instrument was used (by Folmer and Jellinek) to conclude that copper is monovalent in covellite and chalcocite [43]. The valence state o f one is a shift from the previous position o f divalent-copper i n chalcocite, which was previously reported by Jellinek [44].  5  Figure 2 -2. Crystal structure of covellite [40].  6  2.1.2 Leaching Implications of the Structures In the structures of chalcocite and covellite, the top of the valence band is formed by a band (sometimes partly filled) of largely copper 3d orbital character which is able to accommodate variations in metal concentration [45,46]. The leaching of chalcocite by ferric ion media requires successive removal of metal atoms from the structure, to form sequences of non-stoichiometric copper sulfides before second stage covellite (CuS) is formed. A l s o , the sulfur sublattice behaves rather as a framework within which the metal atoms are quite mobile even at ambient temperature. The triangular co-ordination of copper atoms has been found only in a few species of sulfides which are rich in copper, others are bornite, anilite and stromeyerite (AgCuS).  2.1.3 Thermodynamic Considerations The stability domains of sulfide minerals in aqueous media are commonly shown in E h - p H diagrams [47]. Figure 2-3 shows the E h - p H diagram for the C u - S - H 0 system 2  at 25°C. It indicates that both chalcocite and covellite are stable in the water stability-zone (acidic and basic solutions). A t low p H values (acidic media), low oxidizing-potentials between 0.2 and 0.4 volts are required (according to the diagram) to convert these copper sulfides into sulfur species and soluble copper. In neutral or basic solutions, insoluble oxides are formed and low oxidizing potentials are required to decompose these minerals. The stability of the species depends on the equilibrium constant (K) of the reactions involved and this constant is a function of the temperature of reactions. Therefore, a specie which is not stable at 25°C may be stable at higher temperatures, while others may decrease or vanish. In order to construct an E h - p H diagram (at a particular temperature) relative to standard hydrogen electrode (SHE) at 298K, there is need to calculate the free energy for all the species involved at this temperature.  7  ^ 1 1  Cu  1  1  "  2 +  "~ ^  !  i  1  1  1  "  1  ^  1 HS0 " 4  _ ^  0.5  ISO4  1  SO  :^£us^^);^^  ^  ^  ^ - ^  ^  C  U  0  -0.0 LU  —  —.  —  ^  —  ^ —  —.  -0.5  1.0  ;—  Cu —1  -2  1  0  1  r-  H SjHS-  CU  2  1  L_  pH  1  1  8  10  1.  -  )s -Cu2  1  12  Figure 2-3. E h - p H diagram for the C u - S - H 0 system at 25°C, activity o f copper ion 2  at 0.01 mol/L and other ionic species at unit activity. 1.5  n  r  T  8  pH  r  10  12  14  Figure 2-4. E h - p H diagram for the C u - S - H 0 system at 75°C, activity o f copper ion 2  at 0.01 mol/L and other ionic species at unit activity.  8  The free energy at any temperature can be calculated from the following equation;  G Where;  T s~*0  O  G-T-  .  /"IO  298-  9-{T-298)S°  2 9  (2-1)  9&  ( T Tin —— I  (2-2)  V 29oJ  The detailed derivations o f equation 2-1 and equation 2-2 by D i x o n [48] are included i n Appendix 1. For the non-ionic species, the heat capacity can be obtained from the following equation;  C°p  T  298  j T + 298\ a + b\ \ 2 J  c  where a, b and c are the heat capacity functions, which can be obtained from the literature [49].  For the ionic species, the method o f Criss and Cobble [50 and 51] can be used to determine the heat capacity by the following equations;  i g g  S° =S° , m  9  =a  T  + L\S°  (2-4)  298  + zS (H )  (2-5)  +  abs  where a and /3 are the Criss-Cobble heat capacity constants, which are available for 60, 100, 150,200, 250 and 300°C. The values at 75°C can be obtained by interpolation. The values o f free energy (G^g) are available i n the literature [52] for equation 2-1. Based on the above techniques, a high temperature (75°C) stability diagram for the C u - S H 0 was developed and this is shown in Figure 2-4. In comparison to Figure 2-3, the 2  9  9-1-  Cu  z  and H S C y stability -zone decreased while that o f S 0 " decreased. The stability4  zone o f covellite ( p H range) also decreased. The direct oxidation o f the minerals to C u 0 2  9  at p H o f about 2.7 is possible at high temperature and there are slight decreases (by about 20 m V ) in the oxidation potentials required to dissolve the minerals at p H o f about 1.0. The examination o f these E h - p H diagrams (Figure 2-3 and 2-4) reveals that chalcocite and covellite decomposition could occur by about nine reactions.  Reduction: Cu S +  2H  +  Cu S +  H  +  Cu S +  2 e"  2  2  2  2 CuS + 2 H  +  2 CuS +  +  + 2 e"  —"> 2 Cu° + H S  (2-6)  + 2 e"  — > 2 Cu° + HS"  (2-7)  - >  (2-8)  2  2 Cu° +  + 2 e" — >  H + 2e"  —  2  Cu S + H S  (2-9)  C u S + HS"  (2-10)  2  >  S" 2  2  Oxidation: 2 CuS + 4 H 0 2  Cu S + 4 H 0 2  Cu S 2  2  + 4H 0 2  2 CuS + 4 H 0 2  •  —•> C u S 2  2 Cu  2 +  + 8H  +  + 8H  +  + S 0 " + 6 e"  (2-11)  + S 0 " + 10 e"  (2-12)  2  4  2  4  — > 2 C u ° + 8 H + S 0 " + 6 e"  (2-13)  — > Cu S  (2-14)  +  2  4  2  + 7 H + H S 0 " + 6 e" +  4  The reduction routes o f leaching chalcocite and covellite have not been reported in the literature. When chalcocite and covellite are oxidatively leached, none o f the reactions depicted by equations 2-11 to 2-14 is observed. The diagrams also predict that covellite w i l l oxidize to chalcocite by equation 2-11 and 2-14, these reactions occur over geologic time [53].  10  2.1.4 Multiple Stage Dissolution of Chalcocite The dissolution of chalcocite in ferric sulfate solutions is usually reported as occurring in two distinct stages;  First stage:. Cu S 2  +  2Fe  3+  2+  >Cu  2 +  + 2 Fe  >Cu  2 +  + 2Fe  :  + CuS  (2-15)  + S°  (2-16)  Second stage: CuS  +  3+  2Fe'  2 +  The second stage is much slower and much more temperature sensitive than the first. These two equations are convenient, but they are over-simplified representations of what actually takes place. In reality, the first stage consists of a series of steps in which the reaction products may or may not be observed experimentally depending on the rate of transformation of the successive phase. Previous electrochemical studies have revealed with accuracy some phase transformations which can take place under different oxidizing conditions. It was proposed by Cavallotti and Salvago [54] that Cu S first transforms to 2  djurleite (CU1.97S), then to digenite (Cui.gS), and anilite (Cui. S), before transforming to 75  blue-remaining covellite (Cui. S). The transition from blue-remaining covellite to normal 2  covellite is believed [14] to be possible only under conditions of high potential or current density (1400 mV/SHE or 9 A/dm ). 2  From the leaching experiments of chalcocite (Cu S) and digenite (Cu^gS) in ferric 2  sulfate Fe (S0 )3 and ferric chloride FeCl , Moh [55] reported that the non-stoichiometric 2  4  3  covellite (Cu, S) is formed only as an intermediate product prior to the appearance of 2  normal covellite. The long period of time required for this reaction is believed to be responsible for the slow kinetics of second stage leaching. Ignoring the transformation to digenite, Moh proposed the following equation for the overall reaction of the solid phase transformation.  11  Cu S = C u 2  1 + x  S = C u S = S°  (2-17)  In this proposition, x varies from 0.1 to 0.4. The idea that the blue remaining covellite is formed and is responsible for the slow dissolution o f second stage covellite became acceptable because little knowledge was available then, on the mineralogy o f the blueremaining (blaubleiblender) covellite. Whiteside and Goble [56] later classified the reaction products into a series o f non-stoichiometric copper sulfides by using X-ray diffraction-patterns (for the structural changes) and electron-microprobe analyzer (for the chemical changes) to analyze the grains. The following sequence o f transformations was postulated;  C u S - > C u i . 9 7 S - > C u i . S - > Cu,. S 2  8  chalcocite  djurleite  digenite  75  anilite  Cui S — > C u S _ ^ C u ;6  geerite  ] 4  spionkopite  u 2  S—>CuS  yarrowite  covellite  A t ten percent copper extraction, the major phase is supposed to be digenite (theoretically based on the stoichiometry o f the solid), while covellite is the major phase formed at fifty percent copper extraction based on the copper to sulfur atomic ratio o f each phase. It was shown that two or more phases could be present at a particular copper extraction level. A t 3 3 % copper extraction, geerite, spionkopite and yarrowite (with a small percentage o f covellite) were present. The experiments  shed light on the  composition o f the previously classified "blue-remaining" covellite mineral, which are geerite,  spionkopite and yarrowite. The yarrowite phase, and to a lesser  extent,  spionkopite are natural occurring minerals and observed copper sulfide components o f leached heaps or dumps, while geerite occurs very rarely as a natural mineral. The experiment, though without solution control and with a limitation based on the fact that compositions  determined  for selected  individual  grains  do not give  the  gross  representative o f the reaction-product, revealed non-homogenous leaching o f the grains and the high rate o f leaching up to the formation o f second stage covellite. In this work, first stage is referred to as the end o f 50% copper extraction and the formation o f second stage covellite. In addition to the above phases, which are formed prior to second stage covellite, another phase was identified by Whiteside and Goble as copper disulfide  12  ( C u S ) . This was identified from covellite dissolution product diffraction patterns, which 2  resembled that o f the high pressure compound C u S . This has been synthesized only 2  under high pressures and temperatures in excess o f 10 kbars and 300°C respectively from covellite and sulfur [57], and has a cubic ( N i S ) pyrite structure. A t 25°C, it is completely 2  stable only above 8 kbars and it is metastable at 130°C and 1 atmospheric pressure [58]. However, thermodynamic data on this compound are not available i n the literature and little or nothing has been reported about its chemical properties. The natural occurring copper disulfide mineral (villamaninite) has been identified and reported i n the literature [59].  2.2  Polarization Behaviour of Chalcocite and Covellite The anodic dissolution reaction o f a covellite electrode was also observed to be  slower than that o f chalcocite by Cavalotti and Salvago [54]; this was attributed to the greater energy required to break the sulfur sub-lattice formed on the covellite. In a similar electrochemical study by Wright [60], potential sweeps for chalcocite produced a smooth potential-current behaviour, while that o f covellite resulted in a series o f passivation events as shown in Figure 2-5. The curves revealed that the polarization characteristics o f synthetic C u S and C u S are markedly different. Whereas the polarization curve for C u S 2  2  was relatively simple and no passivation effects were apparent, the curve for C u S produced a series o f current peaks indicating that some type o f passivation was taking place. It was probable that the current peak is due to a process o f film formation and subsequent breakdown, or the successive oxidation o f the intermediate reaction products. In order to investigate the influence o f existing reaction layers on the polarization behaviour o f C u S and C u S , a series o f anodic polarization curves were produced for 2  electrodes which had been previously oxidized at specified potential for 24 hours. These are shown in Figure 2-6 and Figure 2-7 [60]. The polarization curves for chalcocite were similar regardless o f the extent or the potential o f the previous oxidation. This suggests that oxidation products have little effect on the rate o f dissolution o f C u S . The slime, 2  13  that oxidation products have little effect on the rate o f dissolution o f C u S . The slime, 2  which formed on the surface o f the electrode was highly porous or non-adherent. In fact, during the pre-oxidation process, the agitated solution became progressively more cloudy as electrolysis proceeded and this could be the reaction products, which were breaking away from the electrode surface. O n the other hand, the pre-oxidized covellite electrode was strongly passivated from the outset, suggesting that the oxidation products formed during pre-oxidation formed a protective layer on the covellite. Notably, the solution remained clear throughout the pre-oxidation process, which revealed that the reaction products did not break away from the electrode surface. However, the oxidation products which could be causing the passivation were not identified and nothing was mentioned about their properties. 2500 T  -3  .  -2  -  1  0  Log i, mA/cm  1  Figure 2-5. Anodic polarization curves for synthetic C u S (1) and C u S (2) electrodes, 2  Scan rate, 1000 m V / m i n (0.01 M C u S 0 , p H 3) after Wright [60]. 4  14  2500  0  -I  -3  -  ,  2  -  ,  1  Log i, mA/cm  0  ,  1 1  2  Figure 2-6. Anodic polarization behaviour of a pre-oxidized Cu S electrode (2) and (3) 2  (0.01 M C u S 0 , pH3) [60]. 4  2000  1500  LU X CO I  > •§-  1000  c o  CL 500  -3  -2  Log i, m A / c m  2  Figure 2-7. Anodic polarization behaviour of a CuS (1) and pre-oxidized CuS (2) (0.01 M C u S 0 , pH 3) [60]. 4  15  Vlao and Peters [17] studied the direct electrorefining of chalcocite (Cu,S) anodes. They found that a volume decrease accompanied the electrolysis, during which pores and cracks formed within the CuS-S° reaction product layer, which were then filled with electrolyte. The copper ions enter the electrolyte at the bottom o f these pores and cracks, and must be transported to the surface by diffusion. Thus, at high currents, the concentration gradient o f C u  2 +  increased continuously until, finally, the solubility of  C u S 0 . 5 H 0 was reached and the salt crystallizes within the pores. A t this point, a 4  2  massive anode would be passivated as shown in curves I and II o f Figure 2-8 [17].  25  2.0  I  M  If  massW* ma* live Parfcutatt  o  I e  j —  1.5  Corr** A/m* 230  300  330  HiSH*  1.0  0.2  1.0  US0« M  0.1  0.3  0.3  r«nf. *C  25  25  23  ioL  1  0.5  2.S  5.0  7.3 Ttaw "™~  It  12.5  HMTI  Figure 2-8: Passivation behaviour of massive and particulate CtijS anodes, after Mao and Peters [17].  16  15.0  However, a particulate anode took much longer to achieve passivation as shown in curve III, because the steep concentration gradients could not be established at the higher porosity. In fact, at the time o f passivation, M a o and Peters calculated that sufficient coulombs had already been passed to convert the anode completely to a mixture o f covellite and sulfur. Hence, they postulated that the passivation i n this case was probably due to the formation o f protective sulfur layer on the remaining covellite particles. It is not obvious i f crystallization o f copper sulfate salt could cause passivation o f grains o f chalcocite or second stage covellite i n acidic ferric leaching o f the minerals.  2.2.1 Effect of Electrolytes on Polarization Venkatachalam and Mailikarjunan [15] carried out polarization tests on anodes o f chalcocite, synthetic copper-iron mattes and commercial 40% grade copper matte i n sulfate, chloride and mixed sulfate-chloride electrolytes (Figures 2-9 to 2-11). A s shown in Figure 2-9, anode passivation occurred much more readily i n the sulfate electrolytes than i n the chloride electrolytes. A l s o , the  addition o f chloride salt to a sulfate  bath delayed  passivation  significantly, even when low concentration o f chloride was added. It was concluded that the addition o f sodium chloride to sulfate solutions reduced the passivity developed at the sulfide electrodes. However, the process through which C l " depassivates the electrode was not provided.  17  30  2 5  c  o  3 20 a  41  •o o c  Current density 0-<K7 A / c m  1 5k  2  o IM C u C t . 2 H 0 2  2  • IM C U S O , . 5 H , 0 10  10  20  30  40 Time, min  60  SO  70  90  T"  ib)  C u r r e n t density 0-031 A / c m 2  40  80 100 Time min  60  o  iMCuCtj^HjO  •  1MCUS0 .5H 0 4  120  2  U0  f  Figure 2-9. Passivation behaviour of chalcocite anodes in cupric sulfate and cupric chloride electrolytes, after Venkatachalam and Mallikarjuan [15].  18  160  3 00,  ,  1  r  Tint* min  Figure 2-10. Passivation behaviour o f C u S anodes in H S 0 and N a C l electrolytes, after 2  2  Venkatachalam and Mallikarjunan 15].  19  4  2-5, Io)  2-0  Currtnt density, 0-047 A / c m  0-51 .2-5  20  20  40  _L 60  T  h  o 100 g/l  NoCl  •  HjSO^  _!  100 g/t  L_  80 100 Tim*, min  I  !  120  U0  T  T  2  16C  lb)  Currtnt density, 0-O31 A/cm  2  o 100 g/l NaCl c  • 100 g/l H SO 2  I 151-  t  240 Tim#,min  Figure 2-11. Effect o f chloride additions on the passivation behaviour of C u S anodes in a Sulfate electrolyte, after Venkatachalam and Mallikarjuan [15]. 2  20  Cavalotti and Salvago [54] investigated the cathodic behaviour o f covellite electrodes i n solutions containing large amounts o f cupric ions and small amounts cuprous ions i n acid solutions o f sulfate, perchlorates, perchloric acid and sulfuric acid which were saturated with H S . When elemental sulfur was present i n the starting 2  electrode, the cathodic process observed was as follows;  S°+ Cu  + 2e  2 +  CuS  (2-18)  When no elemental sulfur was present initially, the potential was found to be near the theoretical value for the following reaction; i  CuS + C u  2 +  + 2 e" — >  Cu S  (2-19)  2  The C u S (in equation 2-19) was identified by using X-ray diffraction patterns. However, 2  i n a cupric chloride C u C l electrolyte, no C u S was obtained, but the following reaction 2  2  occurred instead;  Cu  + Cl" + "  2 +  e  > CuCl  (2-20)  It was postulated that chloride de-passivates covellite through this reaction by removing cupric ions from solution and preventing reaction represented by equation (2-18). F r o m this investigation, (cathodic) reaction products were postulated to be responsible for passivation.  2.2.2 The Effect of Temperature on Polarization The passivation behaviour o f particulate C u S anodes at various temperatures is 2  shown i n Figure 2-12, after M a o and Peters [17]. A t 22°C, they postulated that the solubility o f C u S 0 . 5 H 0 was l o w enough to passivate the highly porous particulate 4  2  21  anode by copper sulfate salt formation. However, at higher temperatures, passivation was attributed to the formation o f a protective sulfur layer.  2.0  •j  o  /  22*C 32*C 40 *C  -1-  ;/  li  11  1.0  0.S  0  1  •  20  40  60  81  Tint — Hours  Figure 2-12. Effect o f temperature on the passivation behaviour o f particulate C u S 2  anodes, after M a o and Peters [17].  2.3  Mixed Potential Theory of Leaching A n electrochemical process which involves two simultaneous reactions has often  been proposed to account for the aqueous oxidation o f sulfides. The reduction o f the oxidant and oxidation o f the mineral take place at the mineral surface, while electrons are transferred through the sulfide lattice. This process is similar to the corrosion o f metals and the electronic conductivity o f copper sulfides makes it possible. The rates o f these  22  electrochemical reactions are functions o f the electrochemical potential in addition to the concentrations o f the species taking part i n the rate-determining process. This theory was originally presented by N i c o l et al. [61] and later refined by Wadsworth [62] and D i x o n [63]. The oxidative dissolution o f chalcocite, and many other sulfide minerals, is a short-circuited electrochemical process, or corrosion couple, involving both anodic and cathodic steps. A n o d i c decomposition o f the mineral phases; Chalcocite:  Cu S  — > C u , . S + 0.2 C u  Digenite:  Cu S  — > CuS  + 0.8 C u  Covellite:  CuS  —> Cu  +  2  L 8  2 +  8  2 +  2 +  S°  + 0.4 e"  (2-21)  +  1.6 e"  (2-22)  +  2 e"  (2-23)  Cathodic reduction o f oxidant: Oxygen:  0  Ferric:  Fe  + 4H  2  3 +  +  + e"  + 4 e" — >  2H 0  (2-24)  Fe  (2-25)  2  —>  2 +  The sulfide mineral, which is a semi-conductor, acts as a conduit for electron transfer between cathodic and anodic sites on the mineral surface. A l s o , because the cell is short-circuited, it develops a mixed potential somewhere between the reversible potentials o f the two half-cells. In general for electrochemical reactions, the difference between the actual cell potential and the reversible potential o f the half-cell, E , is called the overpotential, r). e  This provides the driving force for reaction. Hence, each half-cell reaction bears a certain fixed relationship between rate and potential (typically described by the Butler-Volmer equation). In electrochemical reactions, this rate is manifested as current density, i (or j i n some literature), which has units o f current per unit electrode surface area. In any electrochemical cell, electrons must be conserved. Hence, assuming equal surface areas for the anodic and cathodic processes, then the mixed potential w i l l be that  23  at which the rates o f both half cells are equal and opposite, according to the rate-potential relationships o f each cell. This concept is illustrated by the polarization curves i n Figure 2-13 [62], for the dissolution o f a metal sulfide ( M S ) i n the presence o f a cationic oxidant N  n +  (such as F e  3 +  in acid solution), which has a more positive equilibrium potential. In any electrochemical reaction, the two half-cell reactions are short-circuited and the solid phase acquires a mixed potential (E). It worth noting that E is not equidistant between the reversible (or equilibrium) potentials E j and E , but is instead located such that the partial currents, e  e 2  /j and I (expressed here as current densities, z and i ,) are equal and opposite. The 2  a  c  anodic overvoltage r | is defined as ( E - E , ) and is positive. The cathodic overpotential r | a  e  c  is defined as ( E - E ) and is negative. The electrochemical basis for electron transfer i n e2  simple reversible reactions may be expressed quantitatively by the Butler-Volmer equation, which relates the current density at a solid-solution interface to the established overpotential (for a single electron transfer process).  i  = z F k exp  /3FE  z F k [Cu ]exp  -(l-/3)FE  2+  RT  RT  (2-26)  The first term in the above equation is the partial current density for the anodic (forward) direction o f equation 2-23 and the second term is the partial current density for the cathodic (reverse) direction. F is the Faraday, k represents the rates for the forward and reverse directions, respectively, i n the absence o f potential. (3 is the transfer coefficient. The activity coefficient is assumed to be unity i n this equation.  24  t  POTEHTIAL E ,-Eq«ilibri«» Potential e  £ -EqulUbriia Potenflo! e  E« Mixed Potential  Figure 2-13. Current density potential diagram showing curves for metal sulfide with rest potential E  e l  and redox-ions with rest potential E  e 2  , which formed mixed  potential E [62].  A l s o , the relationship between the rate (current density) and the concentration o f reactants can be quantified by using the simpler forms o f the Butler-Volmer equation. This depends on the type o f leaching. Generally, there are three major types, according to Dixon [63] and these are presented as follows;  25  2.3.1 Type I Leaching This type is observed when the exchange current densities o f the two half cells (ferric/ferrous and the mineral) are similar i n magnitude, but the reversible potentials o f the half-cells are far apart. The situation is shown schematically in Figure 2-14 below. The current density o f either half-cell reaction is j , the subscript d represents dissolution. A  This particular diagram represents the simplest situation to analyze, since both the ferric and the dissolving mineral are i n their so-called "high-field regions" and the mixed potential intersects their Tafel slopes. The anodic and cathodic overvoltage are equal. In some cases, the cathodic overvoltage w i l l be much more than that o f the anodic or vice versa. The anodic dissolution rate o f many sulfide minerals is found to be limited by a single electron transfer step. Hence, the anodic (dissolution) current density, which is similar to the first term i n equation 2-26 may be taken here as:  (2-27)  The cathodic (ferric reduction) current density is:  (2-28)  I f we set these equal and solving for the mixed potential:  RT a  a  +1- a  F  3+  In  26  (2-29)  Hence, the leaching current density is:  JJ  = Ja  = -Jc  In most cases, a  a  =  2  F  k  (2-30)  '  « a = 0.5. Hence: c  1.5  (2-31)  The rate o f leaching is proportional to the square root o f the ferric ion concentration.  tn j.  In j  Figure 2-14. Polarization curve for Type I Leaching.  2.3.2 Type II Leaching This type is observed when the exchange current density o f the oxidizing couple is higher than that o f the dissolution reaction by several orders o f magnitude, and the mixed potential thus corresponds to the reversible potential o f the oxidizing couple as  27  shown in Figure 2-15. For this to occur, the reduction of dissolved oxidants takes place more or less reversibly on the surface o f the minerals. This is always true for the Cu /Cu 2 +  T  couple and often true for the F e / F e 3 +  couple. The mixed potential can be  2 +  represented as follows; •  0  RT F  T  a 2.  (2-32)  Fe  a  Fe-  B y inserting equation 2-32, into equation 2-27 then; exri  a FE a  RT  .3+  *Fk„  Fe  Fe  ,0.5  (2-33) 24  In this type, the leaching rate is proportional to the square root o f the ratio o f concentrations o f the oxidized and reduced forms o f the oxidant.  Figure 2-15. Polarization curve for Type II Leaching.  28  2.3.3 Type III Leaching In some systems, the leaching reaction is particularly fast and the leaching rate is limited by mass transfer of the oxidant to the mineral surface. This is shown in Figure 216 and in this case, the leaching reaction w i l l proceed at the limiting current density o f the ferric/ferrous couple (cathodic limiting current density). Hence, the leaching current density is given by:  h =  FD. ,_dL * ? e  s  3  i  r  .  * ™„[  F e  1'  ]  (2-34)  In this type, the leaching rate is proportional to the oxidant concentration and this w i l l be the case when the reversible potentials o f the half-cell reactions involved in the leaching are relatively far apart. It is expected that Type III w i l l revert to Type II leaching above a certain critical oxidant concentration.  E  Type III Ft*  <->  A A  /'* r  Ml  In k  Figure 2-16. Polarization curve for Type III Leaching.  29  In ]  2.4  Application of Mixed Potential Theory to Leaching The qualitative application o f mixed potential theory to the kinetics o f chalcocite  and covellite leaching is useful i n providing an explanation for the observed sudden change from one kinetic regime to another and could provide a fundamental explanation for the sequential phase transformations associated with some sulfide leaching reactions. In order to apply the mixed potential theory (which has been presented in section 2.3), the first step is to obtain experimentally the polarization curve o f the mineral. Then, calculate the cathodic curve for the ferric/ferrous couple (based on the electrode kinetics o f this couple) w h i c h has been previously investigated on platinum electrodes [64]) and finally, combine the polarization curve o f the mineral with that o f the ferric/ferrous couple.  2.4.1 Application of Mixed Potential Theory to the Leaching of Chalcocite The approach enumerated i n section 2.4 was used by L i et al. [65] to construct the steady-state polarization curves for anodic dissolution o f chalcocite and (theoretical) cathodic reduction o f ferric ions on graphite. This is shown in Figure 2-17; the data used for the anodic polarization o f the chalcocite were obtained from the previous work o f Wadsworth and Zhong [66]. From this figure, the mixed potential o f the half-cell is close to the anodic rest potential (of the mineral) at low concentrations o f ferric ions. A t l o w ferric ion concentration range o f 0.005 and 0.05 m o l / L , Type III leaching is envisioned, the dissolution o f chalcocite w i l l occur at small anodic over-potentials and large cathodic over-potential. Hence, the kinetics are determined principally by the diffusion rate o f ferric ions. A t the cathode, the back reactions which involve ferrous ion may be neglected. A t high concentrations, the anodic over-potential increases and a change o f mechanism to Type I or II may be expected. Though Figure 2-17 does not indicate the possibility o f having the mixed potential very close to the reversible potential o f the ferric/ferrous couple, this has been observed i n some leaching experiments, i n w h i c h the rate depends on the ferric/ferrous ratio.  30  20  Potential (mV) Figure 2-17. Polarization curves for anodic dissolution o f C u S (0.1 m o l / L C u S 0 , 2  4  0.5 m o l / L M g S 0 , 1 mol/L H S 0 ) and cathodic reduction o f ferric ions 4  2  4  on graphite (0.1 mol/L F e , 1 mol/L H S 0 ) at 25°C [65]. 2 +  2  4  2.4.2 Application of Mixed Potential Theory to the Leaching of Covellite The polarization curves for anodic dissolution o f covellite and (theoretical) cathodic reduction of ferric ions on graphite is shown i n Figure 2-18 [67]. The data used for the  anodic polarization o f the  mineral were obtained  from the  experimental  polarization o f a covellite anode in 0.1 m o l / L C u S 0 solution. This figure is similar to 4  31  that o f Figure 2-17, except that the polarization o f the ferric/ferrous couple is shown i n the cathodic direction. The mixed potential o f the half-cells (which is the potential where i= i = -i ) is represented by the voltage o f the dotted lines at different ferric/ferrous a  c  concentrations. A t high concentrations o f ferric (ferric/ferrous curves 1 and 2), the leaching rate is independent o f the ferric concentration as a result o f the limiting dissolution rate o f covellite. The plateau current corresponds to the voltage region where the limiting current density was observed during the polarization o f covellite anode. The electrode reaction did not proceed until the applied current was increased, leading to a very high electrode potential. The limited diffusion rate o f copper was speculated at this plateau, although this was inconclusive. For qualitative application, it is obvious from Figure 2-18 that at high concentrations o f ferric ion (0.1 and 0.25 m o l / L ) , the anodic overvoltage is very large, which indicates type II leaching. The back reactions at the cathode, which involves ferrous ions can not be ignored, though the kinetics are expected to be controlled mainly by the oxidation o f the covellite. A t very l o w concentration o f ferric (0.001 mol/L), there is large cathodic over-potential and the leaching rate is expected to be controlled by the limited rate o f diffusion o f the ferric ions to the covellite. Type III leaching is envisioned. In the middle range o f ferric concentration (0.01 m o l / L ) , there are cathodic and anodic over-potentials and the rate o f leaching is expected to be controlled by both the oxidation kinetics o f covellite and the reduction kinetics o f the ferric. This tends towards Type I leaching. Although, the qualitative information provided by Figure 2-18 on the effect o f ferrous ion concentration is limited, it shows that at 0.1 and 0.01 mol/L ferric concentration, an increase in ferrous concentration w i l l affect the rate o f leaching inversely.  32  F. (0.23M)) ,2+ >  /?  (XO  -2  K)  r.^do" » /f^dO* W  T«*"dO  H  2 +  3  2  r ^ d o " * M) r . ( i o * M) / T « d O " M) F . ^ d O " * M)  f^dO  3  5  Figure 2-18. Polarization curves for anodic dissolution o f CuS at 55°C, 0.1 mol/L C u S 0 , 0.1 mol/L H S 0 and cathodic reduction of ferric ions on graphite. 4  2  4  33  2.5  Chemical Leaching Kinetics Dutrizac and Macdonald [36], in a review paper on ferric leaching o f sulfides,  tabulated the findings o f all previous kinetic studies on chalcocite and covellite. These, along with the results o f more recent studies, are tabulated below i n Table 2-1 and Table 2-1. The first stage leaching o f chalcocite is very fast relative to second stage leaching, or the leaching o f natural covellite. The concurrence is that the rate o f the first stage is controlled by the diffusion o f oxidant to the mineral surface. This view is supported the observed low activation energies (which are between 1.0 and 7 kcal/mole). In most o f the experiments, the rate was found to be first-order dependence on ferric ion concentration, which implied that the rate o f diffusion o f cuprous ion through the lattice is so fast that it never becomes a rate-controlling step. This assumption is corroborated by Jost [68], that copper ion in minerals like chalcocite has high diffusivity, which is as great as that o f aqueous diffusivities, so the cuprous ion diffuses to the reaction site very fast and then react with ferric ions. None o f the previous work considered the depleting effect o f copper ion concentration (which is available for the first stage) on the rate o f reaction. This may explain a change o f rate observed in all the cases especially, when the copper extraction is greater than 25%. The second stage leaching o f chalcocite (or leaching o f covellite) is typically very slow, and the rate is thought to be controlled by the anodic dissolution reaction. This view is supported by high activation energies (between 15 and 25 kcal/mole ) i n the ferric sulfate system, an indication o f chemical reaction control kinetics. The rate dependence on ferric ion concentration decreases at low levels and there is little or no dependence at high levels. This results presented in Table 2-1 and Table 2-2 were obtained i n the absence o f bacteria. A l s o , it bears noting that no attempt was made to control the potential in any o f these experiments. Therefore, since pure ferric solutions were used i n every case, one would have expected the oxidizing potential to fall markedly from time zero in every case, which may explain some o f the discrepancies with regard to dependence o f the rates on ferric ion concentration.  34  Table 2-1. Kinetics o f first stage ferric leaching o f C u S . 2  Material  Medium  Rate dependence on [Fe ] Independent 3+  Natural minerals C u S Ores and Tailings 2  Pure natural Cu S 2  Ferric Sulfate Ferric Sulfate  Activation Energy (kcal/mol)  Independent  Ferric Chloride  Temp. (°C)  25  Low  60-105  Mass transfer  5  1.5  30.-60  6  5.3  60-90  Mass transfer Mixed kinetics Mass transfer  7  Solid state  8  Slight dependence  Low  20-70  Cu S  Ferric Sulfate  Pure synthetic  Ferric Chloride  Independent  0.8  20-80  Cu  2  C u S and 2  10  ~  diffusion o f  Cu S Synthetic  1  ~  Natural 2  Mass transfer  23-95  Ferric Sulfate  2  Ref.  Low  Synthetic Cu S; Rotating disk  First order between 0.005 and 0.05 mol/L  Ratecontrolling process  +  Ferric Sulfate  First order  5-6  5-80  Mass transfer  9  Ferric Sulfate  First order  6.7  28-70  Mass transfer  4  Ferric Sulfate  First order at  2.79  30-90  Mixed kinetics  2  Cu S; Rotating disk u  Mounted natural crystals Natural Cu S 2  low [ F e ] , 3+  decreasing at higher [Fe ^ 3+  35  Table 2-2. Kinetics o f second stage ferric leaching o f C u S and ferric leaching o f C u S . 2  Material  Medium  Activation Energy (kcal/mol)  Temp.  High  35-95  22  T< 60  Chemical  12  8  60-80  Mixed kinetics  12  3+  Natural minerals  Ferric Sulfate  Independent up to Fe >1 g/1  Synthetic  Ferric Sulfate  First order at Fe  Cu S; 2  (°C)  Ratecontrolling process  Ref  Rate dependence on [Fe ]  11  3+  <0.005 mol/L and  Rotating  independent at  disk  [Fe ] > 0.005 3+  mol/L Synthetic Cu S; dispersed particles  Ferric chloride  Increased with increasing F e cone, i n the range 0.25-1.0 mol/L  25  20-80  Chemical  8  Mounted natural crystals  Ferric Sulfate  Increased with cone, i n the range o f 0.0128- 0.424 mol/L F e  14  40-70  Chemisorption / Chemical  4  Independent at  20  30-90  Chemical  3  18  15-95  Chemical  13  18  30-90  Electrochemical  2  2  3 +  3 +  Synthetic powders; Dispersed particles  Ferric Sulfate  [Fe ] >0.005 3+  m o l / L , first order [Fe ]<0.005 3+  mol/L Pure synthetic and natural crystals; Rotating disk Natural Cu S 2  Ferric Sulfate  Independent at [Fe ]>0.005 3+  mol/L, first order at [Fe ]< 0.005 3+  mol/L Ferric Sulfate  ~ one-half order  Initially, all the cuprous ions within the chalcocite matrix are extremely mobile. A s reaction proceeds, approximately half the copper becomes fixed into the second stage  36  covellite phase and is immobile. The following reaction occurs at the digenite/blueremaining covellite interface [2];  (Cu ) S "_^ V +  2  2  where V  C  u  C  u  +  Cu S ' 2 +  2  +  Cu  + ( m )  (2-35)  is a cuprous ion vacancy and (m) denotes a mobile ion. The structure o f  covellite, which is given by equation 2-35, is unstable and changes according to the following reaction;  Cu S " 2 +  2  >l/2 (Cu ) (S ) " +  (2-36)  2  2  2  The mobile cuprous ion i n equation (2-35) is then free to diffuse to the particle surface, where it is oxidized by ferric ion thus;  Fe  3 +  + Cu  +  — ^  Fe  2 +  + Cu  (2-37)  2 +  The second stage kinetics are controlled by an electrochemical reaction, which occurs all over the porous surfaces o f the particle. Marcantonio [2] found the ferric ion dependence to be about 0.3 order, but assumed one-half order. This confirms the shift from a diffusion controlled mechanism  during the first stage o f leaching to  an  electrochemical controlled mechanism during the second stage.  2.6  Effect of Parameters on the Kinetics in Sulfate Media  2.6.1 Effect of Stirring Speed A s long as off-bottom conditions and the limiting thickness o f the laminar layer around the particles are achieved i n the reactor, stirring speed has virtually no effect on the leaching kinetics o f chalcocite and covellite [2].  37  2.6.2 Effect of Temperature The effect o f temperature on the first stage dissolution rate has been reported consistently to be little on the first stage leaching and that reaction proceeds rapidly at room temperature [1]. The various activation energies reported i n the literature are included i n Table 2-1. The kinetics o f the second stage dissolution rate are much slower and more temperature sensitive than that o f the first stage. The reaction rate is extremely slow at room temperature and increases with increase in temperature. The various activation energies obtained for the second stage leaching are provided in Table 2-2. Linear rates were reported for the second stage by the previous investigators, however, a dualmechanism over the temperature range o f 25 and 80°C was reported by Thomas and Ingraham [12]. In their work, the observed rate was controlled by a chemical reaction at temperature below 60°C and above this temperature, controlled by solution transport.  the rate was reported to be  A l s o , there were initial induction periods at a l l  temperatures, but these deviations from other investigations were not explained clearly i n their work and the ion, which was responsible for the mass transfer at high temperature was not mentioned.  2.6.3 Effect of Particle Size The effect o f particle size on the leaching o f chalcocite has been inconsistent. A direct relationship was obtained by Marcantonio [2] when the rates were plotted against reciprocal radius, which did not support his conclusion that rate was controlled by diffusion through the product layer. If diffusion through the product layer was the rate controlling step, the rate would have been proportional to the reciprocal radius squared. The second stage has been reported to be independent o f the particle size [12]. A direct relationship was observed between the rates and the outer surface area o f the disk in the leaching o f covellite disks [12].  38  2.6.4 Effect of Ferric Concentration The effects o f ferric ion concentration on the rate o f leaching have  been  summarized i n Table 2-1 and Table 2-2 respectively. A n independent relationship has been reported between rate and ferric ion concentration [1] in some previous experiments. A l s o i n some investigations, the rate was said to be directly proportional to ferric ion concentration (first order), but at higher concentrations, the dependence decreased [4,6,9]. The second stage response to ferric ion concentration is much less than that o f the first stage. In dilute solution o f ferric ions (which varies between 0.0005 and 0.005 mol/L), a direct relationship was observed by Thomas and Ingraham [12] and at concentration higher than 0.005 mol/L, the rate was almost independent o f the ferric ion concentration. Dutrizac and MacDonald [13] i n a different investigation observed the same relationship, though the ferric ion concentration increased up to 0.3 m o l / L , the rate remained the same. However, L o w e [4] observed an increase i n rate up to 0.42 m o l / L , while Marcantonio [2] obtained a dependence on ferric concentration up to 0.61 m o l / L and a rate order o f 0.3 (which was approximated to 0.5 in his work). One o f the reasons for this inconsistency could be the deviation o f the redox potential i n all the previous experiments.  2.6.5 Effect of Acidity A t constant ferric strength, the rate o f leaching is independent o f the sulfuric acid concentration over the range o f 0.03 mol/L and 0.3 m o l / L [12]. A t a concentration less than 0.03 m o l / L , the rate reduction observed has been attributed to ferric hydrolysis and precipitation. The main purpose o f the acid may be limited to keeping the trivalent iron i n solution. Other investigators [1,2] have found the rate to be relatively independent o f the acid concentration. Nevertheless, it w i l l be important to control acidity especially when other side reactions (which are acid consuming or generating) can occur such as gangue reactions and oxidation o f ferrous to ferric ion.  39  2.7  Effect of Parameters on the Kinetics in Chloride Media In this section, the kinetics o f chalcocite leaching in ferric chloride is presented. A  summary has been included in Table 2-1 and Table 2-2.  2.7.1 Effect of Temperature A s shown in Figure 2-19, the kinetics of second stage leaching o f chalcocite are much more temperature sensitive in chloride media than in sulfate media. A l s o , the activation energy increases as leaching proceeds, due perhaps to the  progressive  formation o f a protective sulfur coating.  Figure 2-19. Effect o f temperature on the leaching o f chalcocite i n chloride ( F e C l 0.5 3  m o l / L , HC1 0.2 mol/L, F e / F e 3 +  2 +  40  10, 2.5 g +150-300 um C u S ) [8]. 2  2.7.2 Effect of Ferric Concentration The rate o f the first stage leaching is first order with respect to ferric ion concentration when the concentration is between 1.25 x 10" and 1.56 x 10" m o l / L . 4  2  However, the rate has no dependence on the concentration i f the level is between 0.25 and 1.0 m o l / L [8]. This suggests that, at a level above a certain ferric ion concentration, rate control shifts from mass transfer o f ferric ions through the stagnant boundary layer to another mechanism. This could be solid state diffusion o f cuprous ions through the chalcocite crystal matrix or desorption (or mass transfer) o f the dissolved products, which 94-  are Fe  9-4-  or C u  away from the mineral surface. However, the solid state process has  been shown by Price [69] to give significantly higher activation energies than are typically reported for the first stage dissolution o f chalcocite. The rate o f the second stage leaching depends on ferric iron concentration with an estimated order o f reaction o f about 0.2 or 0.3.  2.7.3 Effect of Particle Size The particle size only affects the rate o f the first stage o f dissolution, as expected for diffusion controlled kinetics.  2.8  Proposed Mechanisms of Sulfate Leaching A complete review o f the previously proposed mechanisms o f chalcocite leaching  in sulfate media is important, since most o f the reaction steps are similar whether the oxidant is ferric or dissolved oxygen.  41  2.8.1 Cathodically-Controlled Mixed Potential Mechanism Marcantonio [2] expressed the two stages o f chalcocite dissolution with the following anodic half cells:  Stage I:  Cu S  ~ >  CuS +  Cu  Stage II:  CuS  ~~^  Cu  S°  2  +  2 +  2 +  + 2 e"  (2-38)  + 2 e"  (2-39)  Brennet et al. [14] measured the reversible half-cell reduction potentials o f these reactions as E , = 440 to 505 m V ( S H E ) and E = 520 to 570 m V ( S H E ) respectively, 0  2  while Sato [70] reported them as E , ° = 530 m V ( S H E ) and E ° = 591 m V ( S H E ) 2  respectively. The standard half-cell potential for ferric reduction is 771 m V and is represented as follows;  Fe  3 +  + e" — > F e  (2-40)  2 +  3+  Assuming an ideal solution, the  E h o f the  Fe /Fe  2+  couple could vary  (theoretically) from 917 m V at the beginning o f the first stage (I) to 781 m V at the end o f the first stage and the beginning o f the second stage (I, II) according to the above stoichiometry, then to 735 m V at the end o f the second stage (II). In this mechanism, the half-cell potential for blaubleiblender is taken as that for normal covellite. Consequently, i f the mixed potential initially falls below the range o f half-cell potential for stage II (according to Brennet), then elemental sulfur w i l l be unstable in the presence o f cupric ion. This is expected from the diagram in Figure 2-20, since the slow discharge o f ferric ion (cathodic reaction) is said to be the slower electrochemical process, the mixed potential is shifted towards the reversible potential o f the anode (mineral). This means a large cathodic overpotential r\ and small anodic overpotential, c  r| - I f the mixed potential is close to that o f the mineral (as proposed), one would expect a a  drastic fall o f the redox potential from time zero, which is not the case when covellite is  42  leached. The E h profile or the profile for the rate o f ferric discharge was not presented by the author. A t low temperatures, the half-cell potential ranges for stage I and II are distinct, and the processes do not overlap. A t higher temperatures, the ranges overlap to a certain extent, denoted "a"  on the diagram. It is doubtful i f the overlap w i l l take place at a  mixed potential, which is close to the reversible potential of.the chalcocite as shown by Marcantonio [2]. Once second stage covellite is formed, two mixed potentials  are  established and one is closed to the reversible potential o f the chalcocite while the other which is lower i n magnitude is closed to that o f the covellite. It is envisioned that some particles leach at higher potential and others leach at lower mixed potential but faster rate. The two stages leach at these potentials until the first stage leaching finishes and the mixed potential is predominantly determined by the kinetics o f the second stage leaching. The leaching o f second stage covellite continues at this newly established mixed potential. When chalcocite is leached, the redox potential does not fall drastically from time zero because the mixed potential is very close to the reversible potential o f the ferric/ferrous couple. This is what is observed in practice during the leaching o f second stage covellite and natural covellite. One would have expected a small cathodic overvoltage and a large anodic overvoltage i n which the rate controlling step is the electron transfer i n the anodic reaction. The electrochemical model proposed by Marcantonio for the second stage leaching is therefore  faulty because it was based  mechanism.  43  on the cathodically-controlled  • O r  I-  0.9  1 -3+ _ .2+ Fe -he = Fe  0.84-  in  E (V) h  0.7 4-  _ E (n) m  0.6-r—-  Em(Ui)  -in-  LL.  —i-  . i* . „ . „ „ /"STAGE C u + S + 2e » CuS < n 0  Bad ™=--5 IvA  Emd)  r^J^J-I  ^  ^  ffl +  C(,2  +  28  ""  C u  (STAGE I)  2  S  0.4 JFigure 2-20. Illustration o f half-cell and mixed potential variations during ferric leaching of chalcocite: 0.1 mol/L C u S . 0.5mol/L F e , 0.001 mol/L F e " and C u " [2]. 3+  2  2  2  2.8.2 Iron Depassivation Mixed Potential Mechanism A n alternative mixed potential model was derived by M a o and Peters [71] as part o f their study o f the acid pressure leaching in sulfate medium. In their model o f oxygen  44  pressure leaching, the first stage o f leaching is controlled predominantly by the cathodic half-cell  (either  oxygen reduction or ferric ion reduction), which  is similar  to  Marcantonio's model o f leaching. During the second stage, however, covellite is passivated by oxygen, leading to a high mixed potential. However, the anode is depassivated i n the presence o f ferric ions, which leads to higher leaching rates and lower sulfate yield. The observed progress o f the reaction is portrayed on the Evans diagram, which is shown i n Figure 2-21. When leaching commences, oxygen is reduced on a fresh chalcocite surface (curve 1), which is steadily transformed to digenite, thus undergoing about 7 % shrinkage. The outer particle surface is rapidly transformed to digenite, which changes the cathodic process to a combination o f oxygen reduction on digenite (curve 2) and pore-diffusion-limited reduction o f oxygen on unreacted chalcocite at the base o f pores (the dotted vertical section o f curve 1). The first step o f the first stage is concluded, when covellite (CuS) is formed on the outer particle surface. The leaching reaction, which had migrated from point A to point B during the first step, now jumps to point C as oxygen is reduced on the covellite surface. Point C represents an increase in both the exchange current (leaching rate) and the mixed potential. A s the digenite covellite interface area decreases, the sloping part o f the anodic curve rises and point C traces curve 3 to higher mixed potentials and lower rates. The second stage begin when point C migrates to D , at which the covellite is passivated by oxygen in the absence o f ferric ions. The mixed potential rises quickly to 2+  the transpassive point E . A t this point, the potential is well above that o f Fe /Fe  3+  couple  and this potential is also probably sufficient to form sulfate according to the following reaction; CuS  + 4 H 0 —> 2  Cu  2 +  + 8H  +  +  S0 " 2  4  +  8 e"  (2-41)  The observed role o f iron as a catalyst i n this system was explained on the basis that it acts as a surrogate oxidant for oxygen at the mineral surface, whereby the anodic  45  dissolution of covellite may be carried out at the lower mixed potential. This is represented by Point D. where the rate is significantly higher and the potential is low enough to give a low yield o f sulfate. Therefore, stage II leaching is envisioned as some particles leaching at point E and others leaching at Point D , with the proportion o f the latter increasing with increasing iron concentration.  , Caffode ^Sartadan Cures I.OnCUiS!  0i • 4H* +  M~WJ}  2. On Cu. jS caret* (sugg.) 3. On CuS c  1.0  04  I  4  cus*s* • cu * • • at  04  C*S •Qh.jS • 0 JC* • • • 0.4a  0.4  OJ  A • B fta SMp. SBOi 1 c s«cBBd Sfep. sap i 0 SttQ* I LccMnQ (ft E S M i lacNng (Fo  Figure 2-21. Schematic Evans diagram o f C u S pressure leaching, after M a o and 2  Peters [71].  46  2.9  Proposed Mechanism of Chloride Leaching The mechanisms proposed by Fisher et al.[72] for the oxygen pressure leaching o f  chalcocite in chloride and sulfate media are presented for comparison purposes. N o iron was present i n the experiments, the pressure varied between 0.039 and 0.165 M P a , while the temperature varied between 303 and 347 K . The first stage leaching was about seventy times faster in chloride media than the rate in sulfate solution. This large difference was attributed to the formation o f the stable chlorocuprate (I) complex C u C l  (according to equation 2-42), which allows cuprous  3  ions to be extracted from chalcocite without oxidation to cupric at the particle surface. The rate o f leaching in sulfate media is proportional to the pressure during the first stage, with an activation energy which is 31.5 kJ/mol and presumably associated with oxygen adsorption. The rate i n the chloride system is directly proportional to the chloride concentration, with an activation energy which is 22.6 kJ/mol and presumably associated with chloride ion (Cl") mass transfer.  2 C u S + 1/2 0 2  2  + 2H  +  + 6 C f — > 2 CuS + 2 C u C l " + H 0 2  3  2  (2-42)  The second stage (in chloride) is independent o f all the leaching variables except temperature, with an activation energy o f 34.6 kJ/mol. This was attributed to the electron transfer i n the anodic (dissolution) reaction. The second stage leaching i n chloride solution is given by the following equation;  2 C u S + 7/2 0  2  + H 0 2  +  6 C f —> 2 S0 " + 2 CuCl " + 2 H 2  4  47  2  3  +  (2-43)  1  Tlm«. 10  minutts  2  Figure 2-22. Effect of chloride concentration on dissolution, at 303 K , 0.086 M P a oxygen pressure, 0.35 mol/L [H ] and 210 x 177 p m particles [72]. +  48  Figure 2-23. Effect o f temperature on the second stage leaching at 0.5 mol/L C l , 0.086 M P a pressure, 0.35 mol/L H and 210 x 177 um particles [72]. +  49  Figure 2-24 [72] illustrates the first stage leaching process in the sulfate system. The adsorbed oxygen on the surface is reduced by one electron transfer mechanism to water as shown by;  0  + H  2  H0  2  + H  2  H 0  +  -> H 0  +e"  + e" - > H 0  +  2  2  (2-44)  2  (2-45)  + H + e" - > H 0 + O H  (2-46)  +  2  OH + H  2  +  + e"  H 0  ,  2  (2-47)  Figure 2-24. Physico-chemical model for first stage chalcocite dissolution i n the sulfate system [72].  The electrons for oxygen reduction are supplied by the oxidation o f C u to C u +  2  particle surface and S  2 +  at the  2  to S  2  in the particle interior. Cuprous ion diffuses from the  particle interior to the surface where it is oxidized to cupric ion which is released into the solution.  50  Figure 2-25 [72] shows the first stage dissolution process in the chloride system. The rate controlling step is believed to be the diffusion of chloride ion through the solution boundary layer to the particle surface. Cuprous ions diffuse from the particle interior to the surface where they react with chloride ions to form C u C l " , which is then 2  3  released into solution. Oxygen adsorbed on the particle surface is reduced as shown in equation (2-44) to (2-47) by the electrons, which are liberated in the oxidation o f S " to 2  s'. 2  2  Figure 2-25. Physico-chemical model for first stage chalcocite dissolution in the chloride system [72].  Figure 2-26 [72] illustrates the second stage dissolution process in the chloride system. Oxygen adsorbed on the particle surface is reduced to water as shown in equation (2-44) to (2-47). The electrons are supplied by the oxidation o f blue-remaining covellite as shown in Figure 2-26. This reaction occurs as a sequence o f simple electrochemical steps involving one electron transfer. Since the rate does not vary with oxygen pressure, it is postulated that the rate-controlling step is the electron transfer in the anodic reaction.  51  C«S • « M ] 0 * X I * * •  CvCij  1  • SO4* • « N * * Tt  Figure 2-26. Physico-chemical model for first stage chalcocite dissolution in the sulfate system [72].  One thing that Fisher et al. failed to point out is that, in light of the work of Mao and Peters, the anodic dissolution of covellite in second stage must be de-polarized automatically in chloride media by the presence of an active Cu /Cu couple at the 2+  +  mineral surface. This couple is absent in sulfate media, though this may be stable at very high temperature (about 200°C or higher). In the presence of sufficient Cl", the actual leaching reactions would be as follows;  2 CuCl " + l/20 + 2 H 2  3  CuS  +  2  Cu  2+  +  ~ >' r  + 6C1" —>  2 Cu  2+  + 6 Cl" + H 0 2  S°+2CuCl " 2  3  (2-48) (2-49)  In any case, the rate of the second stage leaching is still significantly faster than that of ferric (or iron catalyzed oxygen pressure) leaching under similar circumstances.  52  2.10 Chalcocite Leaching in Other Media Fisher [73] has compared the rates o f chalcocite leaching i n oxygenated solutions. In non-complexing systems (such as S 0 " , N 0 " , C10 "), the rate o f first stage leaching is 2  4  3  4  very slow (second stage leaching was not observed within the duration o f the experiment) at l o w temperature. In complexing systems which require chalcocite oxidation (such as C l " and N H ) , 3  the rate o f first stage leaching is very fast and the rate o f second stage leaching is slow. In complexing systems which do not require chalcocite oxidation (such as CN") there is only one stage o f leaching, and it is extremely fast (probably mass transfer controlled), like the dissolution o f a soluble salt.  2.11 Summary of Literature Review It is widely known that chalcocite oxidation occurs roughly in two distinct stages, the first stage leaching o f chalcocite is very fast relative to the second stage leaching. The rate o f first stage leaching is controlled by the diffusion o f oxidant to the mineral surface. This is supported by the low activation energies (between 1 and 3 kcal/mol) and the firstorder dependence on ferric concentration. O n the other hand, second stage leaching o f chalcocite (or the covellitic phase) is slow and is controlled by the rate o f the chemical reaction. This is supported by the high activation energies (between 15 and 25 kcal/mol) and the small dependence o f the rate on ferric concentration at l o w levels, and very little or no dependence at high levels. It is worth noting that no attempts were made to control the redox potential, which would be expected to fall from time zero i n every case as oxidant is consumed. This may be responsible for the discrepancies previously reported with regard to the dependence o f rates on ferric concentration and other parameters. Hence the objective o f the present work was to control the redox potential i n each experiment i n order to quantify the variables independently. The experimental procedures required to achieve this objective are presented i n the following Chapter.  53  T I M « ,  1 0 minuUt s  ure 2-27. Chalcocite dissolution in various media [73 ].  54  CHAPTER 3 EXPERIMENTAL PROCEDURES The experimental program developed for this study can be divided into three phases; 1. Controlled potential chemical leaching o f chalcocite and second stage covellite; 2.  Bacterial leaching o f chalcocite and second stage covellite;  3.  Mineralogical characterization o f the reaction products to examine the compositional changes which occur during leaching.  3.1.  Chemical Leaching Experiments The chemical leaching o f chalcocite by ferric/ferrous ions under controlled  potential was achieved by using potassium permanganate solution. The resulting phase from the leaching o f chalcocite (second stage covellite) was also investigated  3.1.1 Sample Preparation and Minerals The  samples  of  chalcocite  and  covellite ores  were  obtained  from  the  Mineralogical Research Center Inc., San Jose, U . S . A . These were i n the form o f rocks with average size o f 2 cm. A mechanical splitter was used to select a fraction with a good representation o f the as-received ore samples. This fraction was used in the mineralogical studies and the leaching tests. Qualitative analyses o f a rock section (from the selected fraction) and a powder were made for comparisons o f the particle size to ensure good liberation o f all the phases. Smaller fractions (about 500 g each), which were good representatives o f the selected fraction o f the ore were crushed and ground to achieve 90% coarser than 150 pm. The choice o f using coarser particles for the dissolution tests was to obtain results which could be upgraded easily to plant scale. Prior to screening, about 500 g o f portions were first screened for 3 hours through the # 115 mesh sieve to remove most o f the fines.  55  The plus 115 mesh fraction was screened for 1 hour on a Rotap shaker by using the Tyler sieves # 35 (425 nm aperture size), # 42 (355 um), # 48 (300 um), # 60 (250 Um), # 65 (212 nm), # 80 (180 nm), # 100 (150 nm) and # 115 (125 nm). The major portions (about 90%) o f these particles were in the narrow sized range o f -425+355, 355+300, -300+250, -250+212, -212+180 and -180+150 nm. These particle fractions were stored i n sealed polyethylene bags until they were required, to minimize oxidation o f the mineral surface. A mono-sized fraction (-250+212 nm) was selected (based on section 3.1.4) for all the dissolution tests and other fractions were used to investigate the effect o f particle size. The cumulative particle size distribution determined by sieving is plotted i n Figure 3-1.  30  100 170 240 310 380 450 520 590 660 730 800 870  940  Size,microns Figure 3-1. Cumulative particle size distribution o f the mineral samples as determined by screening with Tyler sieves.  56  3.1.2 Chemical Analysis of the Sample A sample (from the selected fraction) was sent to Chem. M e t Consultants Inc. ( C M C ) and International Plasma Laboratory (IPL) for chemical analysis. A comparison was made between the two results from the laboratories. The results o f the chemical analysis provided guides on the minerals, which were present in the sample before the commencement o f leaching. Analytical results from the two laboratories have little variation (less than 0.2%) except for copper compositions, which shows about 0.5% and 0.7% difference between the two laboratories for chalcocite and covellite respectively. The average gravimetric composition o f each element is summarized in Table 3-1.  Table 3-1. Results o f chemical analyses for the as-received and reground minerals. Sample  s-  Insol.  (%)  (%)  (%)  3.25  21.80  21.70  2.96  (%) 1.51  79.87  0.00  20.13  20.13  0.00  0.00  65.32  3.25  27.80  27.80  2.26  1.37  66.49  0.00  33.51  33.51  0.00  0.00  (%)  Fe (%)  70.48  C u S (theoretical) Covellite  Chalcocite  1  2  1  C u S (theoretical)  Cu  2  2  Others  3  Assay results are average values obtained from duplicate analyses. 2  A c i d insoluble constituent, indicative of siliceous gangue. Other elements such as bismuth, arsenic and antimony.  3.1.3 Mineralogical Characterization of the Sample In order to corroborate the chemical analyses and to obtain the phase abundance, qualitative analyses by X-ray powder diffractometry were carried out. The X - r a y diffraction pattern o f the chalcocite feed was helpful i n determining the initial phases (minerals) present before leaching. The transformation o f these phases could then be monitored by a similar method. The instrument used for this purpose was a Siemens D5000 powder diffractometer, which uses monochromatized CuKoc radiation. This was  57  operated at 40 m A and 40 k V . The spectra were collected from 3 to 70° (26) in 0.01° steps at a count-rate o f 0.4 s per step. The small samples (grains) from the ground and screened fractions were ground further i n a mortar to obtain fine powder. The fine powder was mixed with ethanol to form a paste, which was spread on a glass slide to form thin layers and allowed to dry before inserting into the instrument for analysis. Based on the phases that were identified (by using X-ray diffractometry) and the chemical analyses (Table 3-1), the phase abundance was determined and summarized i n Table 3-2.  Table 3-2. Results o f mineralogical composition o f the chalcocite sample Covellite  Chalcocite  Bornite  Pyrite  Quartz  Others  Chalcocite  (%) O.01  (%) 80.56  (%) 11.48  (%) 3.49  (%) 2.96  1.51  Covellite  89.34  N/E  0.00  7.03  2.26  1.37  Sample  2  1  This may be enargite as revealed in section 4.1. 2  N o t estimated.  3.1.4 Selection of Monosize Particles This qualitative identification o f the phases was used as a litmus test to select the mono-sized fraction used in the leaching tests. In order to select this fraction, a small quantity o f grains from each fraction was spread on different stubs and then carboncoated. These grains were analyzed by scanning electron microscope, which is equipped with an energy-dispersive detector. The elemental composition analysis o f all the phases were made. The size fraction (-250+212 jam) i n which all the previously identifiedphases (section 3.1.3) were present and well liberated was selected for dissolution tests.  58  3.1.5 Chemical Leaching Apparatus and Procedures A  schematic diagram o f the controlled potential leaching set-up (and the  monitoring equipment) is shown i n Figure 3-2. A t the heart o f the apparatus was a stirred tank reactor (STR), which enabled the redox potential i n m V , and other variables such as temperature and p H to be carefully measured and controlled. The S T R had the following features; 1. Air-sealed, baffled (by three plexiglass baffles) glass reactor with a working volume of 1 L ; 2. Overhead variable-speed stirrer; 3. Gas sparge line through which nitrogen gas was added; 4. Reflux condenser to control evaporative solution losses; 5. A p H controller, which delivered acid or base as required; 6. Redox-potential controller coupled to a peristaltic pump, which added potassium permanganate at rate set to regenerate ferric from ferrous ions and thereby maintained a constant redox potential. The interface circuit between the controller and the pump was developed i n house. This was used to ensure proportional control (redox potential only varied by 1 m V ) and data logging o f the redox potentials into a desktop computer; 7. Air-tight plexiglass cover with eight ports, which contained a condenser, sampling frit, nitrogen sparge tube, stainless-steel stirrer shaft, potassium permanganate tube, sulfuric acid replenishment tube, redox potential probe and p H probe; 8. A pitched-blade axial flow impeller was inserted at the end o f the shaft to provide proper mixing and to ensure effective particle suspension at 900 rpm; 9. The entire reaction assembly was clamped into place i n a water bath, which has a temperature controller to maintain constant high temperature operations.  To start an experiment, the water bath was first heated up to the desired temperature. A 1-liter solution (which contains known concentrations o f ferric sulfate F e ( S 0 ) . 5 H 0 and ferrous sulfate F e S 0 . 7 H 0 i n distilled water) was added to the 2  4  3  2  4  59  2  reactor and brought to the operating temperature in the bath. Five grams o f chalcocite from the mono-sized particles o f -250+212 p m were charged into the reactor (in a l l the tests) and the redox potential and p H controllers were initiated. The leach solution redox potential was kept constant by oxidizing the ferrous ions (which were formed as a result o f the leaching reaction) back to ferric ions. This was done with potassium permanganate solution, which was added by a perilstatic pump. The perilstatic pump was controlled proportionally by a redox potential controller, which used inputs from the solution submerged redox potential probe. A platinum redox electrode combined with  a  silver/silver chloride reference electrode was used. In order to maintain the initial acidity o f the solution, a p H controller system, which uses inputs from the pH-probe to initiate mechanically (on/off), a perilstatic pump to add 6 M sulfuric acid into the solution was used. A gel-filled combination p H probe with a silver/silver chloride reference electrode was used i n all the experiments. The standard test for the first stage leaching was at 35°C, ferric ion concentration o f 0.116 m o l / L , ferrous ion concentration o f 0.0202 mol/L and 0.095 m o l / L o f sulfuric acid. The redox potential o f this solution was 501 m V at 25°C (vs. A g / A g C l ) . The same conditions were used as a standard test for the second stage leaching, but at a temperature o f 75°C. The higher temperature (for the second stage leaching) was chosen based on the experimental observations described in section 4.3.1. Solution samples o f about 2 m L were taken from the reactor through a glass frit to keep all solids i n the reactor (during leaching) and these were analyzed for copper by atomic absorption spectrometry. The sampling time varied from 1 to 30 minutes from the beginning to the end o f the test and the residence time varied (from 30 minutes to 72 hrs) depending on the leaching parameters. After leaching for the desired residence time, the reactor contents were suction filtered and the residue was washed thoroughly with deionized water as described in Section 3.1.6.  60  Figure 3 - 2 . Schematic representation of the controlled potential leaching set-up.  3.1.6 Role of Potassium Permanganate The role of potassium permanganate in the chemical leaching experiments was to control the solution potential without it itself taking part in the leaching. The leaching of the minerals was by the ferric/ferrous sulfate solution, which was introduced at the beginning of each test. Since ferrous ion formation from the leaching reaction would reduce the redox potential of the solution, there was a need to oxidize this back to ferric  61  ion and maintain the initial redox potential throughout the duration o f each test. The (acidic) oxidation o f ferrous ion by potassium permanganate is represented by the following reaction:  Mn0 " +  8 H + 5 Fe +  4  > Mn  2 +  + 4 H 0 + 5 Fe  2 +  3 +  2  (3-1)  The standard potential (with reference to hydrogen electrode) o f potassium permanganate in acid solution, E  is 1.51 volts, therefore permanganate i n acid solution is a strong  e  oxidizing agent. In  the  presence  o f manganese(II) ions, the  following  side reaction  with  permanganate, which is called the Guyard reaction [74] could occur;  2 Mn0 " + 3 Mn  > 5 Mn0  +2H 0  2 +  4  2  + 4H  2  (3-2)  +  The free energy change for reaction 3-2 was calculated by Turney [74] as - 61.1 kcal/mol. The precipitation o f the manganese dioxide depends on this free energy change, the solubility o f manganese dioxide and the hydrogen ion concentration. Highly acidic conditions (about  0.1 mol/L H S 0 ) were used i n the volumetric oxidation by 2  4  permanganate to prevent the formation o f manganese dioxide. The mechanism involved in this side reaction has been thoroughly investigated by Tompkins [75] and the following reaction steps are involved.  Mn  +Mn0 "+ H 0 4  HMn0 " + M n 4  2 +  + OH"  2 Mn Mn  4 +  Mn  2 +  Mn  4 +  >  2  3 +  + H 0 2  + OH + OH"  Mn  + HMn0 " 4  + OH"  > 2 Mn0 + H 0 2  >  Mn  >  Mn  >  >  Mn  2 +  3 +  3 +  + Mn +  (3-5)  4 +  + OH  (3-6)  + OH"  Mn(OH)  62  (3-4)  2  + H  3 +  (3-3)  (3-7) > Mn0  2  (3-8)  The manganese dioxide is formed primarily by reaction (3-8) because reaction (3-4) is slow. The addition o f the first hydroxyl ion in equation (3-8) is the ratecontrolling step in the hydrolysis o f the tetravalent manganese ion. The reaction in equation (3-5) is inhibited by the presence o f complexes i n solution. The decomposition o f permanganate i n alkaline solution is represented as follows;  4 M n 0 " + 4 OH"  >  4  4 Mn0 " 2  4  + 2H 0 +0 2  (3-9)  2  Turney [74] calculated the free energy change o f the reaction as -25.0 kcal. A t the saturation solubility o f oxygen in water, the reaction gives the green manganate ion in sodium hydroxide solution (which is above 1 N concentration). When the concentration o f the sodium hydroxide is less than 1 N , the manganate disproportionate according to the following equation;  3 Mn0 " + 4 H 2  >  +  4  2 Mn0 " 4  + 2 H 0 + Mn0 2  2  (3-10)  Other side reactions o f permanganate (dis-proportioning) in strong alkaline solutions could be through the following two consecutive partial reactions;  M n 0 " + e"  >  4  Mn0 "  M n 0 " + 2 H 0 + 2 e" 2  4  2  (3-11)  2  4  >  4 OH" + M n 0  2  (3-12)  Reaction (3-11) is relatively rapid, while reaction (3-12) is slow [76]. The standard potential E o f reaction (3-11) is 0.56 volt and o f reaction (3-12) is 0.60 volt. e  In order to avoid the side reactions represented by equations 3-2 to 3-12 i n the use o f potassium permanganate to maintain the ferric/ferrous ratio, it was necessary to use a highly acidic medium, pure distilled water and almost perfectly pure  potassium  permanganate (reagent grade standardized solution). These conditions were strictly  63  employed in this work. I f ferric ions were to be absent initially, the calculated amount o f potassium permanganate (based on stoichiometric) required to leach 0.056 moles/L o f copper would be 0.0224 moles/L. In this work, ferric/ferrous solutions were used (and present at the commencement) in the leaching experiment to reduce the volume o f potassium permanganate solution required to maintain the redox potential and to ensure that the kinetics were that o f ferric leaching. The volume (or quantity) o f potassium permanganate used i n the experiments varied with the initial ferric concentration, kinetics o f the reaction and the residence time. The maximum volume o f potassium permanganate used i n the experiments was 100 m L (of 0.1 mol/L standard solution). This was equivalent to about 0.012 mol/L at the end o f the experiment. In acidic medium, ferrous oxidation by equation (3-1) can be considered as titration  of  ferrous  ions  by  permanganate  because  it  takes  place  (almost)  stoichiometrically and rapidly. The potential during the titration corresponds to the redox 3+  potential for Fe  2+  /Fe  and rises rapidly. The potential o f the half-cell involved is given  by the following; E g/A ci A  g  =  1-3110-0.0946 p H + 0.01181og [MnQ^l  (3-13)  [Mn ] 2+  B y using solution redox potential o f E  A g  /  A g C  i = 0.501V and p H = 1.1 (which is the 9-t-  standard experimental condition i n this work), the calculated ratio o f M n 0 " to M n 4  is  -27  3.22 x 10  . This low value demonstrates the absence o f free M n C V (which is needed for  the initiation o f the side reaction i n equation 3-3). This calculation was confirmed by a leaching experiment (controlled chemical leaching), in which ferric/ferrous ions were absent and leaching o f the solid minerals was carried out by potassium permanganate solution.  64  3.1.7 Slurry Filtration Method The slurries were filtered using the re-pulp wash method. This method involved filtering the solids twice. After the first filtration, the filtrate volume was recorded and the filtrate solution sampled. Then the solids were washed in a volume o f deionized water which was approximately four times the filtrate volume. The essence o f this repulping step was to wash dissolved copper containing solution out o f the first filter cake. The solids were suspended i n 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 sampled. The filtrate and wash solutions were analyzed for copper and iron concentrations. The final filtration residues were dried for two or three days, at ambient temperature to minimize the tendency for oxidation o f sulfide sulfur or elemental sulfur.  3.1.8 Analytical Methods The soluble iron (ferric plus ferrous ) and the dissolved copper i n the solution were analyzed in-house by atomic absorption spectrophotometry ( A A S ) , while the solid samples were sent to local Vancouver analytical laboratories for analysis. The dissolved ferrous iron was determined by titration with eerie sulfate. The filtration residues were divided into several parts for various analyses. Parts o f the residues were analyzed for copper and iron by A A S after digesting i n aqua regia/bromine. The sulfur species were determined by a sequential leach/digestion procedure. The total sulfur was determined gravimetrically as B a S 0  4  after digesting i n a K B r - B r / H N 0 2  3  solution. The other sulfur  species were determined gravimetrically after carrying out the following procedure. Initially, elemental sulfur was dissolved i n boiling perchloroethylene (the boiling point is 121°C), then the perchloroethylene leach residue was boiled with 10% N a C 0 2  to selectively dissolve sulfate. Finally, the N a C 0 2  Br /HN0 2  3  3  3  solution  leach residue was digested i n a K B r -  solution to dissolve the remaining sulfur, which was assumed to be sulfide.  65  3.1.9 Determination of Copper Extraction and Sulfide Oxidation Copper balances were conducted for some tests to understand the reaction mechanisms and to ensure the calculated copper extractions were reasonably accurate. I f the metal out / metal i n mass ratio did not fall within the target variation o f 0.98 to 1.02, the solution and / or residue analyses were repeated to check for any analytical error, and in some cases the tests were repeated. The mass balance yielded a calculated head assay, C « , , defined by: ca  c  res \  m °l+m  f  Cu,cal  c  Ca  •100%  (3-14)  Where mc°l is the mass o f copper solubilized, m™ is the mass o f copper i n the leach residue and m is the total starting mass o f the head sample. The percent copper extraction t  was based on the calculated head assay:  Percent copper extraction  m,C u  (c^/iooOfo)  •100%  (3-15)  The sulfur species analyses o f the head and leach residue were used to calculate percent sulfide oxidation levels by using the following formulas:  Total S "oxidation (to S° + SO ,") 2  2  ( ™,head _ „ . r e s A '"sulfide '"sulfide head sulfide  m,  S "oxidation to S° = 2  "'sulfur  •100%  (3-16)  J  •100%  (3-17)  V '"sulfide J  S "oxidation to SO4" = (total sulfide oxidation) - (s "oxidation to S°) 2  Where m^  2  de  is the mass o f sulfide sulfur in the head sample, m^  Me  (3-18)  is the mass o f  sulfide in the leach residue and m™ . is the mass o f elemental sulfur in the leach residue. fw  66  3.2  Mineralogical Characterization of Reaction Products The following techniques were used for the qualitative and quantitative analyses  of the reaction products.  3.2.1 Qualitative Analysis The same techniques used for the feed materials in Section 3.1 were utilised for the X-ray powder analysis o f the leached residues. The leached grains were also analyzed by using a scanning electron microscope, which was attached to an energy-dispersive detector for spot analysis. The leached grains (which were washed and dried) were coldpressed into plugs made o f epoxy-resin and polished using alumina slurries down to 1 pm. These were carbon-coated for use in scanning electron microscopy ( S E M ) . The S E M - E D X provided qualitative analysis vis-a-vis the elemental composition analysis o f each grain by using spot analysis. These polished sections o f the particles were used to reveal the phase changes, which occurred during leaching and where the phases were initiated. Subsequent to elemental analysis, different spots and grains were mapped for compositional analysis. A l s o , the  scanning electron microscope ( S E M ) was used to observe  the  dissolution features on the surface o f the particles.  3.2.2 Phase Compositional Analysis The same grains, which were leached under similar conditions and washed in distilled water were analyzed for chemical composition by using an automated C A M E C A S X - 5 0 electron-probe microanalyzer ( E P M A ) . The operating conditions were 12 k V and 1.2 nanoamperes  for the accelerating voltage and beam current respectively. The  standards used were  synthetic covellite for C u , S and pyrite for Fe. Apparent  concentrations were corrected for absorption, secondary fluorescence and atomic number effects by using the Z A F program to determine the analytical compositions o f the grains (or spots), previously mapped in Section 3.2.  67  3.3  Bacterial Leaching of Chalcocite The bacterial experiments  were  similar to the  standard  chemical leaching  experiments except that oxygen and bacterial inoculum were introduced.  3.3.1 Bacterial Culture and Nutrient Media The  bacterial  culture  used  was  a mixed culture  containing Thiobacillus  ferrooxidans. The culture was initially obtained from the Gibraltar M i n e and was used in previous studies at the University o f British Columbia. It had been grown on the Gibraltar concentrate (chalcopyrite) in shake flasks and serial transferred at the exponential growth stage. The shake flask culture contained 16 g o f solid and 100 m L o f "modified 9 K " solution. This is similar to the nutrient solution developed by Silverman and Lundgren [77], but differed in the concentration o f ferrous iron. The modified 9 K had no ferrous ions (equivalent to OK). In order to ensure that a bacterial culture with optimal activity level was used for the bacterial leaching experiments, the culture from the shake flask was grown on the chalcocite feeds i n a 2-L nutrient solution (which differed from that o f the shake flask). This was necessary to adapt the culture to the chalcocite minerals and to have a leach solution with total iron concentration similar to that o f the chemical leaching experiment. The batch culture nutrient solution was the Silverman and Lundgren solution containing 0.116 m o l / L ferrous iron (equivalent to 6 K ) and without potassium chloride (KC1) to avoid chloride dissolution o f the mineral. The batch culture was initiated by transferring the contents o f a flask (at exponential-growth stage) into a 1.5 L nutrient solution, w h i c h contained 24 g o f chalcocite (with particle size o f P  9 0  75 pm). The 2 L volume was then  achieved by adding nutrient solution. The tank and the tank l i d for the batch culture were made from acrylic material. A tank diameter to tank height ratio o f approximately 0.6 was used. The tank was baffled and the slurry stirred by a four blade, forty-five degree pitch down type impeller with a diameter, which is 0.4 times the tank. The stirring speed was the same as that used for the chemical leaching experiments. A i r was sparged into the  68  tank below the bottom impeller. The air supply came from a compressor and was C 0 2  enriched to 1% v/v C 0 using a C 0 gas cylinder and a gas proportioner. A gas flow2  2  meter was positioned at the inlet of the tank; this was used to regulate the C0 -enriched 2  air flow-rate at about 0.25 L air/L pulp/minute. The temperature of the batch culture was controlled at 35°C by using an immersion heater connected to a temperature controller. The pH of the culture was maintained at about pH 1.5 by addition of 6M H S 0 as 2  4  required.  3.3.2 Bacterial Leaching Experiment This was similar to the batch-culturing, but the experiment was done in the glass reactor used for the chemical leaching tests. The particle size was -250+212 pm and the working volume was 1 L. The pulp density was 0.5% (w/w). The experiment was initiated by adding 750 mL of 6K nutrient (with pH of 1.74) into the reactor. This was the Silverman and Lundgren solution containing 0.116 mol/L ferrous iron (equivalent to 6K) and without potassium chloride to avoid chloride effect. The solution was brought to 35°C by the water bath. The reactor was inoculated approximately 2 hrs after starting the experiment, but this was not done until the solid minerals were added. The time interval between the addition of solid and inoculum was 5 minutes. A sample of the inoculum was taken before inoculating the reactor and the leach solution was also sampled immediately after inoculation. Time zero was considered to be the time of inoculation. The inoculum consisted of 160 mL of a stock batch-culture in the late exponential growth phase. This was a solidfreesolution taken after allowing the batch-culture slurry to settle for 3 hours. The initial 1 L working-volume was achieved by adding nutrient and the solution was then sampled. The acidity was maintained at about pH 1.8 by adding some 6M H S 0 2  4  and when the acidity started falling below pH 1.74, a lime slurry was added. The solution pH and redox potential (vs. Ag/AgCl) were measured. The dissolved copper and iron concentrations were analyzed by using AAS. The initial concentration of  69  copper (inoculum-Cu) was subtracted from the subsequent concentrations to calculate copper extraction. A sterile control experiment was conducted in the same reactor with the same conditions, except that 10 m L o f bactericide (2g/L thymol in methanol) was substituted for the bacterial inoculum.  70  CHAPTER 4 R E S U L T S AND DISCUSSION  4.1  Mineralogical Characterization  4.1.1 X-ray and Microscopic Analyses of Feeds Analysis o f the powder by X-ray diffraction showed that the chalcocite sample consisted mainly o f chalcocite, with traces o f bornite, quartz and pyrite. This was used to determine the phase abundance, which was presented in Table 2-2. The complex X-ray patterns o f the natural chalcocite samples are shown i n Figure 4-1. These patterns were collected after purification o f the material and this was used to establish the phases that were present before the commencement o f leaching. These major phases as shown i n Figure 4-1 are chalcocite, pyrite, bornite and quartz (the djurleite phase could be considered an alteration phase o f chalcocite). The S E M - E D X analysis o f a rock section o f chalcocite (~ 2.5 m m diameter) revealed only a chalcocite phase. The spot analysis o f this phase is shown i n Figure 4-2. The K peak intensity o f copper is a  twice that o f the sulfur i n this phase, i n accordance with the atomic ratio (which is 2:1). A compositional image analysis o f a rock section o f covellite revealed principally two phases as shown i n Figure 4-3. There were some chalcocite phases zoned i n the covellite. X - r a y display analysis (obtained by using S E M - E D X spot analysis) o f the covellite phase is shown i n Figure 4-4. This revealed a copper to sulfur atomic ratio o f about 1.0. The K  a  peak o f copper is slightly lower than that o f sulfur because copper has a higher atomic number than sulfur. However, the X-ray powder diffraction analysis o f the covellite powder revealed the presence o f pyrite and quartz i n addition to covellite and chalcocite phases as shown i n Figure 4-5.  71  cs CD CM *H GO Ui cn *H I  "<£i  c 3  1 it  I"  .CO  kO  EO  w <E CO  IS <E K  <r  CO CN  / <E E-  <r  a K  A3 O CO  (Ll E(i,  _i_ 6Z'9ifrZ  _i_  1 00' 0  _1_ 6 Z ' 9 A b Z  sq.unoo  Figure 4-1. X - r a y diffraction pattern o f chalcocite powder.  72  -a  00' 0  - d) 111 C +> +> j : i! J | — (0—+> ^-(--. » 4) 91 -C - fc u 4) fc tt S 3 <- O - P T " '03 fcQCi .  _J  w  £  nj  J r r 3'ifflnli  ,.), cr.11 <yCO*-<•--< 0.1 ©:(T)<T)  W  « 3—3  3  CO  .  .  3  IS •+>  . < • >r-t <H C C •- 3t38 .J - , M S»J  S) (3D Lfl Q O  . 6 3 0  - •U I M Q (S3 > i l r i O f f . i CO >£l 'XI1 j 1 [  LOTHlXllTr  CS S ^ Gl Q i i i i i i i i i i nKN-fOWCO ••rmror-i  In Figure 4-6. the grains from mono-sized particles o f -250+212 pm chalcocite feed showed all the phases, which were previously established by the X-ray diffraction analysis of Figure 4-1. It was intriguing that all the phases were liberated in grains in a rock section (with 2.5 mm diameter), which indicated only one phase. Based on the X-ray (powder) diffraction analysis and the qualitative analysis by S E M - E D X , the mono-sized fraction o f - 2 5 0 +212 p m chalcocite was selected for the leaching experiments. A l s o , the phases o f interest which were selected for analysis were chalcocite, covellite, bornite, pyrite and other non-stoichiometric copper sulfide phases which could have formed during the dissolution reactions.  Figure 4-2. X-ray display (elemental composition) analysis o f a chalcocite rock section.  73  Figure 4-3. Backscattered electron image o f a rock section o f covellite showing chalcocite zoning in covellite.  | K-ray Disp lay 1  —  107 IB rs  • cov*l  s Ca  1 •  2.«  L_  A  .._ 4.0  6.0  8.1  J8,0  JL.  Figure 4-4. X-ray display (elemental composition) analysis o f a covellite rock section.  74  '••!' S (33  (Ig 0 0 ' 0 f c S Z I  IS  S IS r-i ® S i l l  s^unoo  00'0  00'0bSZt  sq.ui-103  Figure 4-5. X-ray diffraction pattern o f natural covellite powder.  75  n m n"j  Figure 4-6. Backscattered electron image o f -250+212 urn grains o f chalcocite showing chalcocite (1), pyrite (2) and bornite (3) at 0% extraction.  4.1.2 X-ray Analyses of Leached Residues The mineralogical characteristics o f the leached residue were compared with that o f the natural covellite mineral. The X-ray (powder) diffraction pattern o f the leached residue, obtained after 44% copper extraction at 35°C by using standard conditions is shown i n Figure 4-7. The major phases in the leached residue were second stage covellite, quartz and pyrite. The skewing o f the peaks in the first stage chalcocite leach residue (Figure 4-7) is apparent and the intensities o f the major peaks in natural covellite (Figure 4-5) are twice that o f the leached residue (especially at 20 = 33°).  76  CO  03  N U  • i£i  X  1T1 'Ln  LD ,  is co CO  'Ln  l'-l r ; cs d.  •ro O Q <E -j ' U i LO Q 'vf ••-< Ln  CN  h S ~ - - ^ i  <E " co  r—' Ji  3  _£ t'j U <E  K  -  B  s:  •••••  N  CO  —* CH Hi H5  ' VJ TO  <r .-.•+•-  <E E->  <r  :•• =i  D O ' a  Q C0  M i' J  tn  ca -f' rH "SI ooSi 5J -JJ a j Ul  ^ 00'SEi<3  00'0  0S'69EE  sq.unoo  00' 0  :  'i'  • JJ  O  Figure 4-7. X-ray diffraction pattern o f the leached residue after 44% extraction, from -250+212 urn grains o f chalcocite at 35°C, 0.116 m o l / L ferric and 0.0202 m o l / L ferrous concentration.  77  pJ (0  It is possible that the presence o f other phases i n the leach-residue was responsible for the skewing o f the peaks, however, covellite was the principal phase. The diffraction properties and the (indigo-blue) color were the same. Several attempts to match the peaks with copper disulfide ( C u S ) , yarrowite ( C u , S ) , spionkopite ( C u , S ) and geerite 2  12  4  (CUJ S ) failed. It is probable that further grinding o f the residue to form powder (which 6  is required by the instrument) altered these phases.  4.1.3 Qualitative Analyses of Leached Residues by SEM-EDX The leached grains were analyzed for elemental composition o f the major phases and local variations i n composition. The morphology o f the leached particles, which could affect dissolution, is shown i n Figure 4-8. Some o f the particles retained their initial sizes (-212 urn), while some were broken into smaller particles (subsequent to cracks and pore formation). A digital image o f a leached grain i n Figure 4-9 reveals the formation o f cracks at the sub-surface and surface. A sponge-like structure enhanced the reaction throughout the porous particle as well as on the surface. During leaching the grains became puffy (like pop corn) and fragile as shown i n Figure 4-10. The cracks and pores were present at both low and high temperature,  the mechanism by which high  temperature produces a higher rate o f leaching (of second stage covellite) must be more than just the physiological changes which are observed on the particle surface. The elemental compositions o f the selected phases in Figure 4-11 are summarized i n Table 41. The grey level intensity shows variations i n the composition o f grains and local variation within the grains.  78  Figure 4-8. Backscattered electron image o f the leached grains o f chalcocite after 30% extraction at 35°C, 0.116 mol/L ferric, 0.020 m o l / L ferrous and 525 m V redox potential (at 35°C). Identified grains in this field are copper-bismuth-sulfide (1) and covellite (2).  Figure 4-9. Digital image o f a leached grain showing cracks and pores after 44% extraction at 35°C, 0.116 mol/L ferric, 0.020 m o l / L ferrous and 525 m V redox potential (at 35°C).  79  Figure 4-10. Backscattered electron image o f a leached grain o f chalcocite, showing puffy and fragile textures after 30% extraction at 35°C, 0.116 m o l / L ferric, 0.020 mol/L ferrous and 525 m V .  Figure 4-11. Backscattered electron image o f selected leached grains for the microprobe (compositional) and elemental analysis o f phases, after 44% extraction at 35°C, 0.116 m o l / L ferric, 0.0202 m o l / L ferrous and 525 m V .  80  Figure 4-12. Backscattered electron image o f selected leached grains after 10% extraction at 35°C, 0.116 mol/L ferric, 0.0202 m o l / L ferrous and 525 m V .  Table 4-1. Summary o f qualitative analyses o f the leached residues by S E M Spot Number  Elemental  Remarks  Composition 1  Fe, S  Pyrite (Figure 4-11)  2  Cu, B i , S  Bismuth, associated previously with chalcocite as a replacement atom for copper (Figure 4-11)  3  Fe, S  Pyrite associated with two other phases within a grain  4  Fe, S  Pyrite  5  C u , S, A s , Fe  Unleached enargite ( C u A s S ) initially present i n the 3  4  (hydrothermal) veins o f the feed as a result o f paragenics effects (Figure 4-11). 6  Cu, S  Covellite; equal intensity o f K peaks  7  Cu, S  Covellite  8  Cu, S  Non-stoichiometric copper sulfide phase; yarrowite  9  Cu, S  Covellite phase forming from yarrowite; this is shown  a  by the grey level difference (between spot 8 and 9 i n Figure 4-11) within the same grain. 10  Cu, S  Non-stoichiometric copper sulfide (yarrowite) single phase. (Figure 4-14 below)  The dissolution features on the yarrowite grains are shown i n Figure 4-14. The C u - B i - S phase was observed as a single phase in one o f the grains and its X - r a y (display) analysis which differed from that o f covellite phase (Figure 4-4) is shown in Figure 4-15. The dissolution features (cracks and pores) which were present on the surface o f this grain revealed the prior leaching o f the Cu-S portion leaving the bismuth and where the C u - B i - S occurred with other phases in a grain, less cracks were formed on this phase unlike the other phases. The origin o f the bismuth is unclear although, substitution o f copper atoms i n the altered chalcocite phase is possible because the chalcocite lattice  82  behaves as a solid solution which can accommodate other atoms. However, naturally occurring copper-bismuth sulfide minerals have not been reported i n the literature.  Figure 4-14. Backscattered electron image o f a yarrowite grain showing cracks.  liBi  Fe 2,0  4.0  Bi  —r8.0  6.0  —I— 10.0  —  r  i4.a I  12,0  kev  Figure 4-15. X-ray analysis o f a C u - B i - S phase, showing the reduction o f K  a  peak  intensity for copper as a result o f some paragenics effects. This differed from that o f Figure 4-4.  83  A s shown i n Figure 4-16, some elemental zoning was observed on the leached particles, such as iron zoning (spot 1) in the copper sulfide grains. This is one o f the paragenics effects, which could have occurred over a geological time as a result o f chemical changes or crystallographic changes to the initial ores. It was apparent that zoning affects leaching because fewer cracks were formed on these zones compared with the iron-free copper sulfide matrix. The preferential leaching o f the iron-free matrix (spot 2) may be due to a galvanic effect making this phase more susceptible to ferric oxidation than the iron zone. The leaching o f each grain is influenced by the adjacent grains, especially when the grains are joined together as shown in Figure 4-17. The pyrite grain (spot 1) w h i c h is adjacent to a non-stoichiometric bornite grain (spot 2). The bornite leached preferentially to the adjacent pyrite phase.  Figure 4-16. Backscattered electron image o f an iron zone (1) in copper sulfide phase (2). The grey level intensity shows differences in composition within the grain. Leach conditions: 35°C, 0.116 mol/L ferric, 0.0202 m o l / L ferrous and 525 m V redox potential (at 35°C).  84  Figure 4-17. Backscattered electron image o f adjacent grains o f bornite (1) and pyrite (2), showing the preferential leaching o f the bornite grain. Leach conditions: 35°C, 0.116 mol/L ferric, 0.0202 m o l / L ferrous ferrous and 525 m V redox potential (at 35°C).  4.1.4 Compositional Changes by Electron Microprobe Analysis The analyzed grains were from the standard leaching experiment carried out at 35°C by using a leach solution o f 0.116 mol/L ferric, 0.0202 m o l / L ferrous and with a redox potential maintained at 525 m V . The initial ferric to copper (in the solid minerals) mole ratio was 2.0. The experiment was terminated after 90 minutes. The chemical analysis o f the leach solution revealed 44% copper extraction at this stage. The chemical compositions o f the phases, which were found i n the leached grains are presented i n Table 4-2. The electron microprobe analyses o f the phases, w h i c h were previously identified by S E M - E D X analysis, revealed that the pyrite phase remained unchanged with the Fe: S atomic ratio remaining at 0.50.  85  Some chalcocite phase (at this copper extraction stage) had changed to yarrowite and covellite with C u : S atomic ratios o f 1.12 and 1.00 respectively. Though the copper extraction was less than 50%, some covellite phases were formed, which means once a phase is formed the transformation to the next phase commences. The bornite phase was transformed to a non stoichiometric phase ( C u F e S ) , while the Fe : S atomic ratio 3 6  4 2  remained approximately at 0.25. This is equal to that o f the initial (stoichiometric) bornite ( C u F e S ) , while the C u : S and C u : Fe atomic ratios decreased. This non-stoichiometric 5  4  bornite is approximately idaite ( C u F e S ) , which has been identified previously (by the 3  4  same technique) i n the leached residue o f bornite at about 40% copper extraction [78].  Table 4-2. Chemical compositions o f the resulting phases by E P M A Phase  W % ( S ) W%(Fe) W % ( C u ) Total  A % ( S ) A%(Fe) A%(Cu)  Fe/S  Cu/S  Pyrite  53.90  46.33  0.10  100.33 66.86  33.00  0.06  0.49  0.00  Pyrite  53.75  46.17  0.09  100.00 66.89  32.99  0.05  0.49  0.00  Covellite  31.90  1.25  66.33  99.48  48.22  1.09  50.60  0.02  1.05  Yarrowite  30.66  1.20  67.84  99.70  46.72  1.05  52.16  0.02  1.12  Yarrowite  30.56  0.10  67.58  98.24  47.16  0.09  52.62  0.00  1.12  Covellite  32.58  0.07  66.60  99.25  49.15  0.06  50.70  0.00  1.03  Idaite  31.75  13.06  54.26  99.07  47.57  11.23  41.02  0.24  0.86  Idaite  31.21  12.92  53.19  97.32  47.60  11.31  40.94  0.24  0.86  W = weight percent; A = atomic percent  The enargite phase ( C u A s S ) which was present initially i n association with 3  4  pyrite either as a replacement deposit or i n the hydrothermal veins o f pyrite (Figure 4-10, spot 5), remained unreacted. The C u : S, C u : A s and A s : S atomic ratio remained at  86  -0.80,  3.00 and 0.25 respectively  (Table  4-3). This  is significant  hydrometallurgical process, which w i l l leach copper from copper  because a  sulfide  minerals  without dissolving arsenic (from the trace mineral) is highly desirable because o f the stringent environmental requirements on arsenic disposal. The composition o f the copper bismuth sulfide after leaching revealed that the chalcocite transformed to covellite, without the dissolution o f the substituted bismuth.  Table 4-3. Chemical compositions o f the minor phases by E P M A Phase  W%(S)  W % ( C u ) W % ( B i ) W % ( A s ) A % ( S ) A % ( C u ) A % ( A s ) A % ( B i ) Total, A%  Enargite  27.72  45.46  0.46  16.07  45.08  37.31  11.18  0.11  93.68  Enargite  28.34  47.03  0.27  18.39  45.50  38.10  12.64  0.07  96.31  Cu(Bi)S  19.50  38.72  42.21  0.00  42.69  42.77  0.00  14.18  99.64  Cu(Bi)S  19.66  37.61  41.38  0.00  43.58  42.08  0.00  14.08  99.74  W = weig it percent; A = atomic percent  87  4.2  Results of Chemical Leaching Experiments: First Stage Leaching The effects o f the leaching parameters on the first stage kinetics are presented in  the following sections.  4.2.1 Effect of Controlled Potential The effect o f controlled potential on the kinetics o f first stage is shown in Figure 4-18. B y allowing the redox-potential to drift from 525 m V (at the beginning o f the experiment) to 460 m V (at the end), the rate o f leaching was strongly altered when compared with the profile which was obtained at constant potential o f 525 m V .  0  2  4  6  8  10  12  14  16  18  20  22  24  26  28  30  Time, minutes Figure 4-18. Effect o f potential (525 m V ) on the first stage leaching of-250+212|im particles at 35°C, 0.116 mol/L ferric and 0.0202 m o l / L ferrous.  88  4.2.2 Effect of Temperature The effect o f temperature on first stage leaching is shown i n Figure 4-19. The results were obtained by using the same leach solution (standard conditions) at different temperatures. The ferric/ferrous ratio was kept constant i n each test. The rate o f leaching was very fast at all temperatures and at 75°C, the first stage was completed within 15 minutes. A n Arrhenius plot based on the reaction rate is shown i n Figure 4-20. The method o f initial rates [79] was used by drawing tangents to the extraction versus time curves. The slope o f the tangent between 0% and 40% extraction was taken as the rate o f leaching i n each experiment. A n activation energy o f 34.0 kJ/mol indicates mixed kinetics, i n which the rate is controlled simultaneously by diffusion and chemical reaction. This is higher than most o f the values (section 2.5) which have been reported previously by about 50%.  0  1  2  3  4  5  6  7  8  9  10  11  12  13  14  Time, Minutes  Figure 4-19. Effect o f temperature on the first stage leaching o f -250+212 p m particles at 0.116 m o l / L ferric and 0.0202 mol/L ferrous.  89  15  -0.7 -I 2.85  2.90  2.95  3.00  3.05  ,  3.10  3.15  3.20  3.25  _l  3.30  1000/T (1/K)  Figure 4-20. Arrhenius plot for the first stage leaching.  4.2.3 Effect of Initial Ferric Concentration The effect o f ferric concentration on the first stage is shown in Figure 4-21. The standard condition was used in all the experiments except that ferrous concentration was changed to 0.0101 m o l / L . The ferric concentrations o f 0.007, 0.0145, 0.029, 0.058, 0.116, 0.174 and 0.232 mol/L correspond to initial ferric : copper (in solid mineral) mole ratios o f 0.125, 0.25, 0.50, 1.00, 2.00, 3.00 and 4.00 respectively. The kinetics were similar at very l o w concentrations (at 0.0145 and 0.029 mol/L) up to 20% copper extraction (which corresponds to geerite) and linear (up to 30% copper extraction) at high concentrations. A t concentration o f 0.058 mol/L, the reaction was markedly more rapid and there may have been a change o f mechanism at this level. A t low concentrations (below 0.058 m o l / L ) , the rate was directly proportional to the ferric concentration (first order with respect to ferric concentration) as shown i n Figure 4-22. In this range, the rate was most  90  likely limited by mass transfer o f ferric to the mineral surface (which is similar to Type III leaching in section 2.2.3) and the reaction rate corresponded to the limiting current density o f the ferric/ferrous couple. A t high concentrations (above 0.058 m o l / L ) , the order o f reaction decreased to 0.5 (as shown in Figure 4-22) with respect to ferric concentration. The change in the order o f reaction at high concentration (as a result o f a change o f mechanism) has not been quantified previously in the literature. 42 40 38 36 34 32 30 28 26  ion,  24 22  o ro  20  X LU  18 16 14 12 -  - * _ F e 3 + (?§ 0.007 M  10  - H - F e 3 + (?I 0.015 M  8  - A - F e 3 + (J § 0.029 M  6  _ ± - F e 3 + (?§ 0.058 M - 0 - F e 3 + (?I 0.116 M  4  - # - F e 3 + (EI 0.174 M  2  - ^ - F e 3 + d I 0.232 M  0 J 8  10  12  14 16 18 T i m e, m i n s  20  22  24  26  28  Figure 4-21. Effect o f ferric on the first stage leaching o f -250+212 p m particles at 35°C and 0.0101 m o l / L ferrous.  91  30  1.30  i  -2.10  i  -1.90  i  -1.70  i  i  -1.50  -1.30  r—  -1.10  n  -0.90  1  -0.70  1—;  -0.50  1  -0.30  I  -0.10  Log Concentration Figure 4-22. L o g rate vs. log ferric concentration plot for the first stage leaching.  4.2.4 Effect of Initial Particle size The geometric mean radius and the area o f the different particle size fractions (based on spherical geometry) are summarized i n Table 4-4.  Table 4-4. Geometric analysis o f the different particle size fractions. Size Fraction (X), urn  M e a n Radius, c m  Area, c m  -355+300  0.01632  0.00335  -300+250  0.01369  0.00236  -250+212  0.01151  0.00167  -180+150  0.00822  0.00085  92  2  The standard conditions were used to obtain the results shown i n Figure 4-23. The instantaneous rates o f reaction were obtained by differentiating the extraction-time data and these were plotted against the reciprocal o f radius squared in Figure 4-24, and against the reciprocal o f radius i n Figure 4-25. A t less than 20% copper extraction, the rate o f reaction was inversely proportional (a straight line through the origin) to the radius squared. This implies that the rate is controlled initially by diffusion o f ferric ions through the product layer to the interface, which is formed by the transformation o f chalcocite, djurleite, digenite and anilite. A t 20% copper extraction (which corresponds to geerite), the rate o f reaction was inversely proportional to the radius. Subsequent to 30% extraction, the rates were equal for the different size fractions as shown in Figure 4-23. The mineralogical studies showed that cracks, breakage and pores were formed before the end o f the first stage, and became much more pronounced at about 2 5 % extraction. The (newly) exposed area reacted at the same rate and the kinetics shifted to chemical reaction control. 45  - f j - X @ -180+150 Micron -250+212 Micron - ± - X @ -300+250 Micron - * - x @ -355+300 Micron  0  2  4  6  8  10  12  14  16  18  20  22  24  26  28  30  Time, minutes  Figure 4-23. Effect o f particle size on the first stage leaching at 35°C, 0.116 m o l / L ferric and 0.0202 m o l / L ferrous.  93  18  0  2,000  4,000  6,000  8,000  10,000  12,000  14,000  16,000  r (cm ) 2  2  Figure 4-24. Rate vs. the reciprocal o f radius squared plot for first stage leaching. 4 ,  0  20  40  60  80  100  r" (cm ) 1  1  Figure 4-25. Rate vs. the reciprocal radius plot for first stage leaching.  94  120  140  4.2.5 Effect of Initial Ferrous Concentration The effect o f ferrous concentration on the kinetics o f first stage leaching is shown in Figure 4-26 and Figure 4-27. The two sets o f results were obtained at l o w and high concentration o f ferric (0.116 and 0.232 mol/L respectively) to corroborate the previous observations in section 4.2.2. Other operating conditions were the same i n each set. A t the l o w concentration o f ferric (0.116 mol/L), an increase i n ferrous concentration up to 0.108 m o l / L had no effect on the kinetics o f first stage leaching (Figure 4-26). This confirmed that the rate o f the first stage is controlled by the mass transfer o f ferric ions to the mineral surface and that the reaction is occurring at the cathodic limiting current density. The back reaction involving ferrous ions at the cathode has no effect on the rate.  Time, minutes  Figure 4-26. Effect o f ferrous on the first stage leaching o f -250+212pm particles at 35°C and 0.116 mol/L ferric.  95  However, at the higher concentration o f ferric (0.232 m o l / L ) , i n w h i c h the mixedpotential becomes more positive and corresponds to the reversible potential o f the ferric/ferrous couple, an inverse relationship between rate and ferrous concentration was observed (Figure 4-27). The fractional order o f reaction was obtained by plotting log rate vs. log ferrous concentration as shown in Figure 4-28. The rates were obtained by the same method used in Section 4.2.2, but the slopes o f the tangents were evaluated between 20% and 50% extraction. The fractional order decreased from -0.15 to -0.4 (approximately) as the ferrous concentration increased as shown Figure 4-29. This is similar to type II leaching, in which an electron transfer mechanism in the anodic (dissolution) reaction is the ratecontrolling step.  0  2  4  6  8  10  12  14  16  18  20  22  Time, minutes  Figure 4-27. Effect o f ferrous on the first stage leaching o f -250+212pm particles at 35°C and 0.232 mol/L ferric.  96  24 26  1.00  SLOPE=-0.15  0.80  0.60  SL0PE= - 0.36  u re i_ X LU  ^ 0.40 O CSI  ra  a> +-*  ra  K  0.20  D)  0.00 -2.50  -2.00  -1.50  -1.00  Log Concentration  -0.50  0.00  Figure 4-28. L o g rate vs. log ferrous concentration plot for the first stage leaching o f -250+212 jam particles at 35°C and 0.232 mol/L ferric. o.oo  -0.10  -0.20  -0.30 fl>  u -0.40  -0.50 0.00  0.20  0.40  0.60  0.80  Concentration, mol/L  Figure 4-29. Reaction order vs. ferrous concentration plot for the first stage leaching.  97  1 .00  4.3  Results of Chemical Leaching Experiments: Second Stage Leaching The product o f first stage leaching is covellite, with a structure and chemical  composition identical to natural covellite. The effects o f leaching parameters on the kinetics o f second stage leaching are presented in the following sections.  4.3.1 Effect of Controlled Potential The effect o f controlled potential on the kinetics o f second stage leaching is shown in Figure 4-30. The ferric/ferrous solutions were the same i n the experiments. B y allowing the redox-potential to drift from 567 m V to 500 m V (at 75°C) for 15 minutes before engaging the controller to maintain the potential at this level for the subsequent period o f time, the leaching profile was strongly altered.  20 10  o i  ,  0  1  , 2  3  Time, hours  4  5  Figure 4-30. Effect o f potentials (500 and 567 m V ) on the second stage leaching o f the -250+212pm particles at 75°C, 0.116 m o l / L ferric and 0.020 m o l / L ferrous.  98  4.3.2 Effect of Temperature The effect o f temperature on second stage leaching is shown i n Figure 4-31. The standard conditions were used at different temperatures and at a constant ferric/ferrous ratio. The redox potential o f the leach solution at 35°C, 55°C, 65°C and 75°C are 525, 545, 556 and 567mV. The second stage was much more temperature sensitive than the first stage. A t high temperature (75°C), the first stage leaching was complete within 15 minutes. The rate o f leaching at 75°C compared favourably with that o f oxygen pressure leaching experiments [71] some o f which were done at higher temperatures.  I 0  1  !  2  '  '  1  1  1  4  1  1  1  6  1  !  1  8  1  1  !  1  1! 10  Time, hours Figure 4-31. Effect o f temperature on the second stage leaching o f -250+212 urn particles at 75°C, 0.116 mol/L ferric and 0.0202 m o l / L ferrous.  99  1  1  (  12  A t 35°C, leaching was very slow and substantial copper extraction was not obtained after 12 hours. A t 55°C, a point o f inflection was observed between 50 and 60% copper extraction, which may indicate (probably) a change o f mechanism or the formation o f an intermediate product (with a corresponding change o f rate). This observation was confirmed by repeating the tests with exactly the same results. A t temperatures higher than 55°C, the point o f inflection was absent, which may indicate rapid decomposition o f the probable intermediate product. A n Arrhenius plot for second stage leaching is shown i n Figure 4-32. The method o f initial rates was used by drawing tangents to the extraction versus time curves. The slope o f the tangent between 50% and 100% extraction was taken as the rate o f leaching in each experiment. A n activation energy o f 97.0 kJ/mol was obtained, indicating that the rate is controlled by chemical reaction.  2.00  i  •  ,  -0.50 J 2.82  2.92  3.02 1/T *1000(1/K)  Figure 4-32. Arrhenius plot for the second stage leaching.  100  3.12  3.22  4.3.3 Effect of Initial Ferric Concentration The effect o f ferric concentration on second stage leaching is shown i n Figure 433. The standard conditions were used in all o f the experiments, except that ferrous concentration was changed to 0.0101 mol/L. Ferric concentrations o f 0.058, 0.116, 0.232 and 0.348 mol/L correspond to initial ferric : copper (in solid mineral) mole ratios o f 1.00, 2.00, 4.00 and 6.00, respectively. Generally, the change o f rate with ferric concentration was less significant when compared to the first stage. A reaction order o f 0.5 with respect to ferric concentration was obtained over the whole ferric concentration range studied as shown i n Figure 4-34. This indicates an electrochemical mechanism which could be either Type I or Type II depending on the effect o f ferrous concentration.  0.0  0.5  1.0  1.5  Time, hours Figure 4-33. Effect o f initial ferric on the second stage leaching o f -250+212 u m particles at 75°C and 0.0101 mol/L ferrous.  101  2.0  1.80  -2.00  -1.50  -1.00  -0.50  0.00  Log Concentration  Figure 4-34. L o g rate vs. log ferric concentration plot for the second stage leaching.  4.3.4 Effect of Initial Particle size The effect o f the particle size on second stage leaching is shown i n Figure 4-35. These tests were run at the standard conditions. There was no relationship between particle size and the rate o f extraction, although one would have expected an inverse relationship between rate and particle size because o f the chemical controlled kinetics obtained i n section 4.3.2. From the mineralogical analysis (in section 4.1), dissolution features (such as cracks and pores) and particle breakage were observed. This resulted i n an enlarged surface area such that the initial surface area o f the particles became insignificant when compared with the new surface area formed.  102  0.0  0.5  1.0  1.5  2.0  2.5  3.0  3.5  4.0  Time, hours  Figure 4-35. Effect o f particle size on the second stage leaching at 75°C, 0.116 m o l / L ferric and 0.0202 mol/L ferrous.  4.3.5 Effect of Initial Ferrous Concentration The effect o f initial ferrous concentration on second stage leaching is shown i n Figure 4-36. The ferric concentration (0.116 mol/L) and temperature (75°C) were held constant i n all o f the experiments. The fractional order o f reaction was obtained by plotting the logarithm o f rate against that o f ferrous concentration as shown in Figure 437. This is a case o f type II leaching (section 2.3.2), in which the rate is controlled by an electron transfer mechanism in the anodic (dissolution) reaction. It is expected that the kinetics w i l l depend on the ferric/ferrous ratio.  103  0.0  0.5  1.5  1.0  Time, hours  2.0  2.5  3.0  Figure 4-36. Effect o f ferrous on the second stage leaching o f -250+212 u m particles at 75°C and 0.116 mol/L ferric.  The effect o f ferrous concentration on second stage leaching at l o w temperature is shown i n Figure 4-38. The ferric concentration (0.116 mol/L) and temperature (35°C) were held constant. The redox potential o f the solutions at 25°C, were 601, 551, 501 and 451mV. These were i n the same range (0.001 to 0.101 mol/L) as the high temperature experiments (75°C). The effect o f ferrous at low temperature was significant up to 70% extraction and subsequently the rates were almost the same at different levels o f ferrous. A t high concentration o f ferrous (0.101 mol/L and 451 m V ) , the reaction did not proceed beyond 70% within 70 hours. A t 75°C, the change o f rate was not observed i n any o f the experiments and reaction proceeded beyond 70% extraction at 0.101 m o l / L ferrous.  104  1.80  1.20 1-10 ! -2.500  1  ,  ,  ,  (  -2.000  -1.500  -1.000  -0.500  0.000  Log Concentration  Figure 4-37. L o g rate vs. log ferrous concentration plot for the second stage leaching o f +250-212 um particles at 75°C and 0.116 m o l / L ferric.  The fractional order o f reaction was obtained by plotting the logarithm o f rate against that o f ferrous concentration as shown i n Figure 4-39. The rates were determined by drawing tangents to the extraction versus time curves. The slope o f the tangent between 50% and 70% extraction was taken as the rate o f leaching i n each experiment. Although, the rate is controlled by an electron transfer mechanism i n the anodic reaction at both high and l o w temperatures, it is apparent that the leaching system o f second stage covellite is closer to type II leaching (section 2.3.2) at high temperatures than at l o w temperatures. A t l o w temperature,  the Tafel slope o f the anodic reaction (mineral  dissolution) changes rapidly as reaction proceeds such that the limiting current density o f the mineral is reached at about 70% extraction and the subsequent leaching occurs at the limiting current density o f the mineral. The driving force for the leaching reaction is the anodic overvoltage and this must be sufficiently large to force the reaction to proceed  105  beyond 70% extraction. The increase in ferric/ferrous ratio may not translate to a significant change i n rate because the dissolution is occurring at the limiting current density o f the mineral. The high ferric/ferrous ratio is still required to force the leaching beyond 70% extraction. A t high temperature, the initial exchange current density o f the second stage covellite is larger than that of the l o w temperature and as leaching progresses the dissolution current density is always higher than the exchange current change o f the covellite. Though the Tafel slope o f the anodic reaction (mineral dissolution) changes as reaction proceeds, the limiting current density o f the mineral is not reached at about 70% extraction and leaching occurs throughout at the dissolution current density which is larger than the limiting current density o f the mineral.  106  0.00 -0.05 -0.10 -0.15 -0.20 •*-»  *  o o  -0.25  -0.30 -0.35 -0.40 -0.45 -0.50 -3.50  -3.00  -2.50  -2.00  -1.50  Log Concentration  -1.00  -0.50  0.00  Figure 4-39. L o g rate vs. log ferrous concentration plot for the second stage leaching o f -250+212 u m particles at 35°C and 0.116 mol/L ferric.  4.3.6 Effect of Ferric/Ferrous Ratio The effect o f ferric/ferrous ratio on second stage leaching is shown in Figure 4-40. The  total iron concentration (0.136 mol/L), total sulfate concentration (0.287) and  temperature (75°C) were held constant i n all o f the experiments. Three different ferric/ferrous ratios were investigated: 0.116/0.020, 0.10/0.036 and 0.082/0.054 m o l / L corresponding to ferric / ferrous ratios o f 5.80, 2.78 and 1.52 respectively. The redox potentials (vs. A g / A g C l ) o f the solutions were 567, 546 and 531 m V at 75°C respectively. The fractional order o f reaction was obtained by plotting the logarithm o f rate against that of ferric/ferrous ratio as shown in Figure 4-41. The rates were determined by drawing  107  tangents to the extraction versus time curves. The slope o f the tangent between 50% and 100% extraction was taken as the rate o f leaching i n each experiment. The second stage leaching is sensitive to the ferric / ferrous ratio at high temperature. In this range o f ferric /ferrous ratio, the ferric/ferrous couple is near equilibrium, with large exchange current density and the mixed potential is within the anodic (second stage covellite) Tafel region. This is a case o f type II leaching (section 2.3.2), i n which the rate is controlled by an electron transfer mechanism in the anodic (dissolution) reaction.  Figure 4-40. Effect o f ferric/ferrous ratio on the second stage leaching of-250+212 urn particles at 75°C, 0.136 mol/L total iron and 0.287 m o l / L sulfate.  108  1.60  1.55  1.50  a  ra * 1.45 a o  _i  1.40  1.35  1.30 0.00  0.10  0.20  0.30  0.40 0.50 Log [Fe ]/[Fe 1 34  0.60  0.70  0.80  2  Figure 4-41. L o g rate vs. log of ferric/ferrous ratio for the second stage leaching at 75°C.  4.4  Effect of Leaching Parameters on Sulfur Distribution The effect o f the redox potential on the reaction product distribution is presented  in Table 4-5. The ferric concentration (0.116 mol/L) and temperature (35°C) were held constant in all the experiments. A t low temperature (35°C) and at the standard conditions (E  A g / A g C 1  =501 m V at 25°C), first-stage leaching (up 40% copper-extraction) occurred  with negligible sulfur formation (2% total sulfide oxidation). A t 35°C, the initial oxidation o f second stage covellite (which corresponds to 5070% copper extraction) was much more favourable at higher potential (601 m V ) than at lower potential (501 m V ) , resulting in a higher rate o f extraction at 601 m V .  109  Table 4-5. Effect o f redox potential on sulfur distribution in the leach residue. Test conditions: 35°C, 0.055 mol/L initial copper, 1.085 g o f sulfur (in the solid mineral) and 0.116 mol/L ferric (leach solution). Redox  Copper  Potential  Extraction  S °  ScuS  Ss04  s  %  %  %  %  ( m V at  Cu/S  Cu/S0  Molar  Molar  Ratio  Ratio  1  4  Time  Cu/S  Hrs  Moles Ratio  1  3  25°C) 1.5  1.00  8.69  22  0.83  0.81  6.06  20  0.90  61.71  1.48  6.69  79  0.76  25.73  50.87  1.87  5.11  72  0.73  30.12  44.50  1.92  4.85  69  0.73  501  40  2.00  0.00  98.00  N/A  501  60  16.24  5.60  78.16  1.00  601  60  19.92  8.04  72.04  451  71  23.02  15.27  501  77  23.40  601  80  25.38  2  N/A  2  Obtained by using the moles of copper which dissolved during the second stage. Not applicable because there was no leaching of copper from second stage covellite at this stage. 'Obtained from the moles of copper remaining in the solid residue.  T w o possible competing reaction paths which may be responsible for the different rates and reaction products (up to 70% extraction) are:  CuS  >  CuS  >  Cu  +  2 +  1/2 C u S  2  S° + 2 e" +  1/2 C u  2 +  (4-1) + e"  (4-2)  The free energy change o f the reaction 4-1 at 25°C is 118.9 kJ/mol. The free energy o f formation o f C u S was estimated by adding the difference between the free 2  energy o f formation o f pyrite (FeS ) and pyrrhotite (FeS) to that o f C u S . Based on the 2  110  estimated free energy o f formation o f C u S and that o f C u S , the free energy change o f 2  reaction (4-2) was estimated to be 26.2 kJ/mol at 25°C. The first path would be more favourable than the second at higher potentials. A n y leaching situation i n which reaction 4-1 occurs more rapidly than equation 4-2 would show a higher initial rate o f extraction and a higher content o f elemental sulfur. A t 60% extraction (which corresponds to 10% extraction from second stage covellite) the elemental sulfur formed at 601 m V was more than that at 501 m V . The corresponding dissolved copper to sulfur mole ratio were 0.81 and 1.00 at 601 and 501 m V , respectively. The elemental sulfur formed at this level o f copper extraction (60% extraction) was 78%) and 69% o f the entire sulfur formed when leaching was prolonged (to obtain copper extraction greater than 70%) at 601 m V and 501 m V respectively. This shows that a substantial part o f the entire sulfur was formed at this stage v i a reaction (4-  !)• The sulfate formed at 60% copper extraction was 27% and 22% o f the entire sulfate formed when leaching was prolonged (to obtain copper extraction greater than 70%) at 601 m V and 501 m V , respectively. A l s o , additional sulfur was formed subsequent to 60% copper extraction, but its increase was less than that o f sulfate. It is possible that the following reactions may be responsible for the formation o f additional sulfur and sulfate:  1/2 C u S  —>  1/2 C u  2 +  + S° +  1/2 C u S + 4 H 0 — >  1/2 C u  2 +  + 8H  2  2  2  +  e"  (4-3)  + S 0 " + 7 e" 2  4  (4-4)  The free energy changes o f reactions (4-3) and (4-4) were estimated to be 92.7 kJ/mol and 296.9 kJ/mol at 25°C. A t about 80% copper extraction (which corresponds to 30% extraction from second stage covellite), the sulfate formed at 601 m V was more than that at 501 m V . The corresponding dissolved copper to sulfate mole ratios were 4.85 and 5.11 at 601 and 501 m V , respectively.  Ill  The effect o f temperature on the product distribution is presented in Table 4-6. The operating conditions were the same i n all the experiments except temperature and time. The quantity o f elemental sulfur formed increased with temperature possibly because the dissolution o f the C u S through reaction 4-3 is more favourable than reaction 2  4-4 at high temperature.  Table 4-6. Effect o f temperature on the sulfur distribution i n the leach residue. Test conditions: 0.055 mol/L initial C u , 1.085 g o f sulfur (in the solid mineral), 0.116 mol/L ferric and 0.020 mol/L ferrous. Leach  Copper  Temp.  Extraction  O  0  %  Ss04  SfJuS  Cu/S°  Cu/S0  %  %  Molar  Molar  Ratio  Ratio  °C  %  35  60  16.24  5.60  78.16  1.00  8.69  55  63  22.82  2.94  74.24  0.92  75  87  56.55  5.91  37.54  1.07  1  4  Time  Cu/S  Hrs  Moles Ratio  1  22 '  0.83  21.51  5  0.81  30.73  4.5  0.56  C u S  2  Obtained from the moles of copper formed during the second stage. Obtained from the moles of copper remaining in the solid residue.  A t the highest temperature (75°C), the oxidation o f copper disulfide through equation 4-3 was very fast such that the copper disulfide from reaction 4-2 was almost instantaneously oxidized. This occurred as i f equation (4-2) and (4-3) were combined into one reaction. The net free energy change o f reactions (4-2) and (4-3) is 118.9 kJ/mol, however the activation energy observed in section 4.3.2 is 97 kJ/mol. The elemental sulfur was 90% o f the total sulfide oxidized. It is possible that the sulfur formed at high temperatures is porous and non-adherent to the mineral surface. Otherwise, one would have expected sulfur to insulate the mineral surface, having an adverse effect on the kinetics.  112  4.5  Bacterial Leach Experiment The effect o f the bacteria is shown in Figure 4-42. The bacterial leaching rates  compared favourably with the chemical leaching rates (at 35°C), for which the same total iron concentration (0.116 mol/L) was used. The initial concentration o f copper and ferric ion (from the inoculum) was 0.0155 and 0.0172 mol/L respectively i n the bacterial leaching experiment. The initial redox potential was 366 m V (vs. A g / A g C l ) . A t high population and activity o f the bacteria, the lag period before the exponential growth phase was greatly reduced as shown in Figure 4-44. A c i d addition was only required during the first stage leaching, while lime slurry was added after 45 hrs (during the second stage). The instantaneous rate o f reaction at different level o f extraction was plotted for bacterial and chemical leaching in Figure 4-43. The significant rates o f extraction were obtained from the bacterial leaching when the redox potential was i n the range o f 450 and 650 m V . The chemical leaching was carried out at constant redox potential o f 601 m V . The rates o f extraction for the chemical leaching were higher than that o f bacterial, but subsequent to 65% extraction, the difference in rates tapered off and became equal at 80% extraction. There was no substantial copper extraction (beyond the acid soluble copper) i n the sterile experiment (without bacteria). Though conclusions cannot be drawn on the mechanisms involved i n this bacterial leaching experiment, the results do reveal how efficient the bacteria are i n maintaining the solution potential and sustaining the rate o f reaction. The rate o f second stage leaching depends on the redox potential (Section 4.3.), which appears to be the primary role o f the bacteria. The substantial addition o f the acid occurred during the first stage, when the redox potential was in the range o f 366 and 512 m V . When compared with the results o f the chemical leaching in section 4.2, the probable reaction path can be represented as follows.  2 F e S 0 + 1/2 0 4  Cu S 2  +  2  + H S0 2  Fe (S0 ) 2  4  b i o l o g i c a l 4  chemical 3  >  Fe (S0 )  >  CuS + C u S 0 + 2 F e S 0  113  2  4  3  + H 0  (4-5)  2  4  4  (4-6)  In this mechanism, there is ferrous oxidation by the bacteria, which increases the ferric concentration. The redox potential was low during the first stage leaching because o f the high consumption requirement o f ferric ions by the rapid dissolution reaction o f chalcocite. Under active bacterial leaching conditions, during which ferrous is oxidized to ferric faster than ferric is reduced by the mineral, high ferric / ferrous ratios and redox potentials (exceeding 650 m V ) are obtained. This was observed during the second stage leaching and the rate-determining step in this (indirect) mechanism is the anodic (dissolution) reaction at the mineral surface. In second stage leaching, the probable path is represented as follows:  CuS  +  Fe (S0 ) 2  4  c h e m i c a l 3  > CuS0 + 2 FeS0 + 4  4  S°  (4-7)  Another reaction which occurred during the second stage was sulfur oxidation (catalyzed by bacteria) to produce sulfuric acid. This was responsible for the decrease i n p H after the previous upward trend.  S°  +  3/2 0  2  + H 0  h i n l n s i r a l  2  >  H S0 2  4  (4-8)  There was a change o f rate at about 70% copper extraction i n bacterial and chemical leaching but this was more pronounced in chemical leaching. A t this stage, the p H o f the bacterial leaching was low as a result o f reaction (4-8). This could mean that sulfur oxidation by the bacteria is desirable for achieving a high rate o f oxidation during the second stage leaching.  114  gure 4-42. Effect o f bacteria on the leaching (first and second stage) of-250+212 p m particles at 35°C and 0.116 mol/L total iron. The initial concentration o f ferric was 0.0172 m o l / L . The redox potential for the chemical leaching experiment was controlled at 601 m V .  115  1.40  1.20  Rate-Bacterial leaching Rate-Chemical leaching  1.00  0.80  0.60  0.40  ro 0.20  0.00 40  45  50  55  60  65  70  75  80  85  90  Extraction, %  Figure 4-43. Rate vs. extraction plots for the bacterial and chemical leaching (second stage) o f -250+212 jam particles at 35°C and 0.116 m o l / L total iron. 750 -,  1  3.0  2.9 2.8 2.7 2.6 2.5 2.4  > E  2.3  CL Ct  2.2  o  2.1 2.0 1.9 1.8 1.7 1.6 1.5 10  20  30  40  50  60  70  80  Time, hours  Figure 4-44. Redox potential and p H profiles of bacterial and chemical leaching.  116  CHAPTER 5 THEORY AND MODELING  5.1  Physico-Chemical Model of First Stage Leaching The leaching rate of chalcocite shifted twice during the experiments (section  4.2.4), first at about 20% copper extraction and second at about 38% copper extraction. Subsequent to 38% copper extraction, the rate decreases drastically. Two steps are obvious and these are described as the first-step which produces up to 20% copper extraction and the second-step, which produces the remaining 30% extraction.  Figure 5-1. Physico-Chemical model for the first step o f first stage leaching.  117  The first step is illustrated in Figure 5-1. In this, x increases sequentially from 0 to 0. 4 and the kinetics o f leaching are controlled simultaneously by the following:  1. Diffusion o f the F e  ions through the product layer to the interface,  3 +  2. Diffusion o f mobile C u from the chalcocite (interior) to the interface, +  3. The oxidation o f the C u to C u , which releases electrons for the ferric reduction, 4. The reduction o f the F e 7+  5. The diffusion o f C u  3 +  ions,  and Fe  7+  through the product layer to the bulk solution.  The particle is transformed sequentially from chalcocite to djurleite, digenite, anilite and geerite ( C u , S ) . The oxidation o f S " to S " also supplies electrons for the 2  6  2  2  reduction o f ferric ions and this can take place in the interior o f the chalcocite particle, or at the interface. Subsequent to 20% extraction, the cracks become enhanced and the second step commences. The leaching o f the second step is illustrated below in Figure 5-2.  Figure 5-2. Physico-Chemical model for the second step o f first stage leaching.  118  Based on microscopic examination (section 4.1), cracks and pores formed on the surface and subsurface as shown i n Figure 5-2. In this model, y increases sequentially from 0.4 to 1.0. The pronounced dissolution features (cracks and breakage) minimized the mass transfer limitation o f ferric ions to the interface and that o f cupric ions to the bulk solution. One would have expected an increase i n the rate up to 50% extraction as a result of increase i n surface area associated with cracks and particle breakage, but this was not the case. This is because the rate o f the first stage (generally) is proportional to the concentration o f cuprous ions which are available (and mobile) for oxidation during the first stage. Though, the dissolution features  became pronounced as the leaching  proceeded the concentration o f the mobile cuprous ions was already at l o w level. The relationship between the rate and the concentration o f cuprous ions is shown i n Figure 53. X is the fraction o f copper extracted. The trendline intercepted the concentration at 0.034 m o l / L , which corresponds to about 39% copper extraction. 0.06  y = 3.7408X-0.1274 R = 0.9892  0.05  2  0.04  0.03  & 0.02 ra 0.01  0.00 0.033  0.035  0.037  0.039  0.041  0.043  0.045  0.047  0.049  0.051  Cu (1-X), mol/L 0  Figure 5-3. Relationship between rate and cuprous ion concentration for the first stage leaching o f -250+212 urn particles at 35°C, 0.116 m o l / L ferric ion and 0.0101 mol/L ferrous ion concentration.  119  5.2  Physico-Chemical Model of Second Stage Leaching The particles break up at the end o f first stage leaching, while the second stage  covellite phase (and negligible elemental sulfur) is formed. Each particle then leaches discretely. The reduction o f ferric ions takes place on the surface o f covellite and the electrons are produced by the oxidation of covellite as shown i n Figure 5-4. This is similar to that o f Peters and M a o [71]. Since a high activation energy is involved in the electron transfers, simple sequential electrochemical reactions involving one electron transfer (in each reaction) are proposed, and the rate controlling step is one o f the electron transfer steps in the anodic reaction. This is clearly illustrated by an electrochemical model.  Cathodic : 2 F e Anodic  3 +  + 2 e"  : 1/2 ( C u ) S +  2  2  > 2 Fe >  Cu  2 +  2 +  + S° + 2 e"  Figure 5-4. Physico-Chemical model for second stage leaching.  120  5.3  Electrochemical Model of Second Stage Leaching A t high temperature (75°C), the second stage leaching commences before the end  of first stage once second stage covellite is formed. The first stage leaching is completed within 15 minutes. Then the mixed potential is determined principally by the reversible potentials o f the second stage covellite and the ferric / ferrous couple. The position o f the mixed potential relative to that o f the  ferric/ferrous  couple can be  determined  (qualitatively) by using the volume o f potassium permanganate required to keep the potential o f the solution constant at the initial stage o f reaction (Appendix 3). The volume required for the first stage was much more than that required during second stage leaching. The electrochemical model illustrated i n Figure 5-5 explains the half order dependence on ferric concentration which was observed in this work (4.3.3). This can be Type I or Type II leaching (section 2.3) and the exact kinetics can be quantified by electrochemical oxidation experiments. Over the range o f ferric concentration used i n this work (0.058 to 0.348 mol/L), the leaching rate may be controlled by both the oxidation kinetics o f second stage covellite and the reduction rate o f ferric ions. If the mechanism is Type I leaching, there would be cathodic and anodic overvoltages and the half order dependence would be on ferric concentration alone, and not on ferric / ferrous ratio. The (calculated) exchange current density o f the ferric/ferrous couple on a graphite electrode (for 0.1 mol/L ferric and 0.01 m o l / L ferrous) is about 9.486 mA/cm  [67], which is significantly higher than the experimental value for covellite,  which is about 1.4 m A / c m [67]. However, high temperature could increase the exchange current density o f the mineral [17] so that it becomes similar i n magnitude to that o f the ferric / ferrous couple and the mixed potential [ E ^ ] then intersects the Tafel slopes o f the two half-cells as anticipated in Figure 5-5. Though the mixed potential is lower than that o f Type II, the rate o f reaction which corresponds to L o g y is higher in this case. The cathodic overpotential is large, the back reaction involving ferrous ions at the cathode is not significant and the rate depends only on the ferric concentration. The rate o f reaction is controlled by charge transfer steps in both the cathodic and anodic reactions.  121  If the mechanism is type II leaching, there would be a large anodic overpotential (with negligible cathodic overpotential) and the dissolution current density would be less than the exchange current density o f the ferric/ferrous couple. Since the mixed potential [E  M(1I)  ] corresponds to the reversible potential o f the ferric / ferrous couple, the rate o f  leaching depends inversely on the ferrous concentration.  Logj Figure 5-5. Evans diagram o f theoretical polarization curves during ferric sulfate leaching o f second stage covellite.  In this work, fractional order dependence on both ferric and ferrous concentrations was observed at low temperature (35°C) and at high temperature (75°C) which indicates that the increase i n temperature does not change the leaching to type I leaching as envisaged i n Figure 5-5. A l s o , the mixed potential which is a kinetic measurement does not change significantly. The rate o f leaching is controlled principally by charge transfer in the anodic reactions and, hence, Type II leaching in which the rate depends on ferric and ferrous is inferred.  122  5.4  Rate Expression for First Stage Leaching The rate o f first stage leaching was fast at low ferric concentration (0.007-0.116  mol/L) and this can be predicted by a mixed kinetics model in which the rate is controlled simultaneously by the diffusion o f ferric ions and subsequent chemical reaction with cuprous ions. The mechanism changes to chemical reaction control at about 20% copper extraction, when the particles break and the cracks become much more pronounced. The rate also follows the concentration o f cuprous ion, which is available and mobile for chemical reaction (Figure 2-3). Initially, all o f the cuprous ions i n the chalcocite are very mobile, but as reaction continues, some o f these become fixed into the resulting phases (solid products) and immobile for further first stage reaction. Hence, a shrinking sphere model, which incorporates the effect o f the cuprous ion concentration profile (section 5.1) is derived to describe the rate o f first stage leaching at l o w ferric concentrations. This is based on the particle leaching theory presented by Fogler [79] and Levenspiel [80]. The rate o f first stage leaching can be entirely chemical reaction controlled, i n which case the rate o f reaction is described by the following expression;  da,  k [Fe ] 3+  =  r  s  dt  (  5  1  )  a  a  A l s o , the rate can be entirely diffusion controlled, in which case the rate o f reaction would be expressed:  da,  k ([Fe ) -[Fe l) 3+  =  c  dt  3+  h  a„  Since [Fe ] i s not easily measured (as is the bulk concentration [Fe +  s  defined i n terms o f  [Fe ] . 3+  b  123  ] ), it must be b  In order to develop a mixed kinetics model, equations (5-1) and (5-2) were equated and solved for the surface concentration o f ferric ions:  [  F  e  3  +  M^f!!k  ]  ( 5  .  3 )  K+K A l s o , there is a relationship between rate and the concentration o f copper i n the solid mineral as explained in section 5.1. This can be represented as follows:  ^  = £[Cu] (l-a,)  (5-4)  0  dt  B y substituting equation (5-3) into equation (5-1) and combining with equation (5-4);  da,  T h e n  W[Cu ] (l-a,)[Fe ] +  =  3 +  0  dt  b  2n a (k+K) 0  a  Where aj = fraction o f copper extracted which can be leached i n the first stage and  a this is cc, = —•; a = fraction o f total copper extracted a = upper limit o f first stage, which is 0.44 a  t  = the leach time (minutes).  k = mass-transfer coefficient (cm/min) c  k = chemical reaction rate constant (in units which w i l l give the rate expression r  units o f min" ). 1  k = rate constant with respect to cuprous concentration, L/(mol-min) n = initial total moles o f copper i n chalcocite sample, m o l / L . 0  [Cu ] (1 -a ) +  Q  x  = concentration o f mobile cuprous ions at different fractions o f  copper extraction.  124  For small particles with negligible shear stress at the fluid boundary, the mass transfer correlation (Frossling equation [79]) for flow around a spherical particle is reduced to the following equation: Sh = - U L  D  =  2  (5-6)  Then,  (5-7)  Where Sh = Sherwood number D - diffusivity (cm /min) 2  d = diameter o f particle i n cm p  The substitution o f equation (5-7) into equation (5-5) produces:  da, dt  Jc k[Cn] (l-a,)[Fe \ 3+  r  0  kA 2n a (l + - ^ ) n  0  fl  This can be simplified further as:  da, _ ^ [ C u ] q - g ) [ F e ] +  3+  o  2 n  1  0« ( O  1  +  b  -y)  Where D* = diameter at which the resistances to mass transfer and reaction rate are equal, that is D =  125  5.5  Rate Expression for Second Stage Leaching The rate o f second stage leaching can be predicted by an electrochemical reaction  model with fractional dependence on both ferric and ferrous concentrations. A t high temperature, the predominant anodic reaction is the oxidation o f covellite to copper and sulfur as discussed i n section 4.4 by the following overall reaction: Cu S 2  <  2  k  >2Cu  a  + 2S°+4e"  2 +  (5-10)  A n d the cathodic half-cell reaction is: Fe  + e" < - ^ - > F e  3 +  (5-11)  2 +  Though two electrons are involved per mole o f copper i o n produced from equation (5-10), an assumption o f a series o f steps involving the transfer o f only one electron is considered. I f only one o f these is rate-controlling, the Butler-Volmer expression may be invoked to describe the net anodic and cathodic current densities, i  a  and i , respectively: c  i« = a ^ ^ a Z  e X  P  Ml  -z F£ [Cu a  a  2 +  ]exp  RT  i  = z F k [Fe ]exp 2+  c  c  (5-12)  RT  c  Ml  z Fk [¥e c  ]exp  c  -<\-P.)FE  RT  RT  (5-13)  A t mixed potential, i = - i , and by assuming f3 = fi « (3; the two equations (5a  c  a  a  12 and 5-13) can be combined to give the following expression:  ^[FE/RT]  = A.z.MCun  A z MFO  +  e  c  A z ^ +A zJc[F a  a  a  c  e  2 +  126  ]  (5-14)  Where E is the mixed potential. In the second stage, there is a large anodic overpotential; hence the anodic back reaction can be ignored i n equation (5-14) and, since the rate is controlled by charge transfer i n the anodic reaction, the rate o f oxidation can be written thus:  dn dt  C u S  _  A  i zF a  F a  =  E  (5-15)  exp — v RT  A  A„z  A k  A zA+A zA[7V ] +  a  c  The ferric/ferrous couple is near equilibrium, with large exchange current density, and the mixed potential is within the anodic (covellite) Tafel region. Thus equation (5-15) becomes:  dn,  ^  dt  r  = Kh  r  [Fe f  V  3+  k)  (5-16)  [Fe Y 2+  K  In terms o f the rate o f copper extraction, we obtain the following:  da dt  da  2  ~dt~  1 dn CuS 2n  (5-17)  dt  0  1  da  « b ~ « a  d  (5-18)  t  Substituting equation (5-16) and (5-17) into equation (5-18)  da.  1  dt  2(a -a )n b  a  AX 0  Fe  (5-19)  Fe  127  The fractional order o f reaction obtained i n this work with respect to ferric / ferrous ratio was 0.42. This fraction can be substituted i n equation (5-19) to obtain the following expression.  0.42  2 dt  da  1 =  A  2(a -a )n b  a  0.42  jC [Fe )_ 2+  0  (5-20)  Where Oj = fraction o f copper extracted which can be leached i n the second stage and this is Oj  a-a.  a = fraction o f total copper extracted a = upper limit o f the second stage, which is 1.00 b  t = time in hrs. k = cathodic rate constant c  h = anodic rate constant in the forward direction a  n = initial total moles o f copper i n the chalcocite mineral 0  The surface area changes with the reaction time as leaching progresses and more pores are formed. The net result o f physical changes to the ore particle w h i c h occur as leaching progresses is called the topochemical effect (Appendix 4). This can be correlated by some common mechanisms [80] such as linear leaching with chemical reaction at the particle surface rate-controlling, and parabolic leaching with the diffusion o f a reactant solute through a porous product layer rate-controlling. In Figure 5-6, the logarithm o f the instantaneous rate o f extraction (of copper available for leaching i n the second stage) is plotted against the fraction unreacted (l-a ). The best power law fit obtained i n this work 2  leads (approximately) to a chemical controlled mechanism with shrinking cylinder geometry.  128  1.60  S L O P E = 0.50 1.40  1.20  fi n  L O P E = 0.51 1.00  oi o  Log Rate-75° C Log Rate-65° C Linear (Log Rate-65° C) Linear (Log Rate-75° C)  • • — —  0.80  0.60  0.40 0.200  0.400  0.600  0.800  1.000  1.200  1.400  1.600  1.800  Log (1-X)  Figure 5-6. L o g rate vs. extraction for the second stage leaching of-250+212 p m particles at 65°C and 75°C, 0.116 mol/L ferric and 0.0202 m o l / L ferrous.  Hence, the anodic surface area can be represented as follows;  A =<pA [\-a f  (5-21)  5  a  0  2  Where <p is a geometric factor and A is the original surface area o f the second stage 0  covellite. Substituting equation (5-21) in equation (5-20) gives:  0.42  da. dt  ^A [l-a ]° £  ~h  5  2(a -a )n b  a  0  2  a  K_  0  0.42  [ [ ^  3  +  ] l  [[Fe ]\  (5-22)  2+  In we introduce an Arrhenius function for temperature, then equation (5-22) becomes;  2 dt  -E/RT  da  =kc  [l-a r  (5-23)  2  129  0.42  Where k =  (pKK  2(a -«aK b  0.42  ~K  (5-24)  K.  0.42  k — Jc* [Fe' ] +  [Fe ] 3+  Then,  ^ - = te dt  Also,  c =ke~  2  E / R T  [l-a f 2  1  =ke  E , R T  0.42  (5-25)  [Fe \ 2+  91QQQIRT  (5-26)  B y using the experimental activation energy which is 97000 J/mole; A t 7 5 ° C , c = 2.7532 x l O Then,  - 1 5  (5-27)  ^  = c\\ - a, 1°  (5-28)  5  dt  L  2 J  B y integration, the following relationship was obtained from equation (5-28);  a =\  '2-cO  2  (5-29)  2  A t 75°C and different ferric/ferrous ratio, attempts were made to fit the model with the experimental data in Figure 5-6. A t 75°C, 0.116 mol/L ferric and 0.0202 mol/L ferrous with ferric / ferrous ratio o f 5.8. c = 0.7943 k =2.885 x 10  14  it, = 1.3788 x 10 14  A t 75°C, 0.082 m o l / L ferric and 0.054 mol/L ferrous with ferric / ferrous ratio o f 1.52. c =0.4526 k = 1.6439 x 10  14  *, = 1.3788 x 10  14  The comparison o f the predicted and actual data is presented i n Figure 5-7.  130  Time, hrs  Figure 5-7. Predicted vs. actual extraction for the second stage leaching o f -250+212 u m particles at 75°C and different ferric / ferrous ratio.  In Figure 5-8, the logarithm o f the instantaneous rate o f extraction (of copper available for leaching i n the second stage) is plotted against the fraction unreacted  (l-a ) 2  at 55°C. The best power law fit obtained in this work leads (approximately) to a chemical controlled mechanism with shrinking cylinder geometry. A t 55°C and by using equation 5-26, c = 3.5657 x 10" £ 16  A l s o equation 5-28 can be used to predict the extraction as shown i n Figure 5-9. A t 55°C, 0.116 m o l / L ferric and 0.0202 mol/L ferrous. c =0.1029 k = 2.8849 x 10  14  k = 1.3788 x 10 14 x  131  (5-30)  0.85 0.80 0.75 0.70 0.65 0.60 (0  C£  O)  0.55  o _l  0.50 0.45 0.40  Figure 5-8. L o g rate vs. log o f extraction for the second stage leaching of-250+212 u m particles at 55°C, 0.116 mol/L ferric and 0.0202 m o l / L ferrous. 1.00  •D  2  Pred. T@55°C  o  •  X  0)  c o  0.20  2  0.10 0.00  4  Time, hrs  Act.  6  T@55°C  10  12  Figure 5-9. Predicted vs. actual extraction for the second stage leaching of-250+212 u m particles at 55°C, 0.116 mol/L ferric and 0.0202 m o l / L ferrous.  132  A t 35°C and by using equation 5-26, c = 3.5394 x 10 A:  (5-31)  _17  A l s o equation 5-28 can be used to predict the extraction at 35°C and at different concentration o f ferrous as shown i n Figure 5-10. A t 35°C, 0.116 m o l / L ferric and 0.020 mol/L ferrous. c =0.0102 k = 2.8849 x 10  14  k, = 1.3788 x 10  14  1.00 0.90 0.80 0.70 0.60 -  0.50  2  0.40  4-1  X O  Pred. Fe2+@0.020M  1 °-  A  30  2  Act.  Fe2+@0.020M  Pred. Fe2+@0.101M •  0.20  Act.  Fe2+@0.101M  0.10 0.00  20  40  60  80  Time, hrs  Figure 5-10. Predicted vs. actual extraction for the second stage leaching of-250+212 um particles at 35°C, 0.116 mol/L ferric and different levels o f ferrous.  In order to test the robustness o f this model, the initial experimental conditions o f Marcantonio [2] were tested as shown in Figure 5-11. It is important to recall that in the work o f Marcantonio, the initial ferric / ferrous ratio was allowed to deviate from time zero, before the commencement o f the second stage leaching but i n Figure 5-11, the  133  initial conditions were used to predict extraction. This may be responsible for the disparity between his data and that o f the model in the present work.  Fe3+/Fe2+ <§ 1.2 1 «  Marcantonio. Fe3+/Fe2+ @ 1.2  0.20  0.10  o.oo Time, hrs  Figure 5-11. Predicted vs. actual extraction (Marcantonio's data [2]) for the second stage leaching o f chalcocite particles at 75°C, 0.24 ferric and 0.2 m o l / L ferrous.  134  CHAPTER 6 CONCLUSIONS The following are the major conclusions reached from this study o f the leaching kinetics o f chalcocite and second stage covellite.  1. The first stage leaching is very fast and second stage leaching is slow at l o w temperatures. Non-stoichiometric copper sulfides are formed before the formation o f second stage covellite and the rate o f reaction o f all these phases are fast enough such that none o f them affects the rate o f second stage leaching. Hence, even though the rate decreases rapidly at about 40% copper extraction, blue-like covellite (Cuj S ) is 2  not formed and second stage leaching does not involve this compound. 2. The first stage leaching is controlled simultaneously by the diffusion o f ferric ions and the chemical reaction. The particles break before the end o f the first stage, thus rendering the surface more reactive, but this does not enhance the rate because o f the low concentration o f mobile cuprous ions (within the crystal lattice) at this stage. A t low ferric concentrations, the rate o f reaction is rapid and controlled by mass transfer of the ferric ions to the reaction surface. The rate dependence on ferric concentration decreases with increasing ferric concentration during the first stage leaching. The leaching mechanism changed at high ferric concentration to an  electrochemical  mechanism i n which the rate is controlled by charge transfer in the anodic reaction. 3. The second stage leaching is controlled by chemical reaction and particle breakage during the first stage enhances the dissolution o f the second stage covellite by providing fresh and large surfaces for chemical reaction. The half-order dependence of rate on ferric concentration and the fractional order dependence on the ferric / ferrous ratio during second stage leaching suggests an electrochemical mechanism i n which the rate is controlled by charge transfer in the anodic reaction. This mechanism operates at both l o w and high temperatures. Although an increase in temperature  135  enhances the rate o f reaction, it does not alter significantly the (electrochemical) mechanism o f anodic dissolution. 4. The rate o f second stage leaching is governed by the redox potential. There is also a fractional order inverse dependence  o f rate on the ferrous concentration. This  indicates that the reversible reaction which involves the ferrous  ions o f the  ferric/ferrous couple is significant in the electrochemical mechanism. Hence, there is a need to have low concentration o f ferrous during the tank and heap leaching, and this is the primary role o f the bacteria. A n active bacterial condition, i n which the oxidation o f ferrous ions takes place faster than the reduction o f ferric ions by the second stage covellite is necessary for process efficiency at l o w temperature. 5. The reaction products o f the second stage are sulfur and sulfate at l o w temperatures, substantial part o f the elemental sulfur is formed first and its accumulation at about 70% copper extraction reduces the leaching rate at low temperatures. The sulfur oxidation by bacteria is desirable for achieving a high rate o f reaction during l o w temperatures leaching o f the mineral. The sulfate is formed substantially during the latter part o f the reaction at low temperature. A t high temperatures, the principal product is elemental sulfur, which is porous and does not inhibit leaching. Another possible intermediate product is copper disulfide ( C u S ) which decomposes rapidly at 2  high temperatures to copper and sulfur. 6. The electrochemical leaching model developed to predict second stage leaching has been validated with experimental data from this work and previous work. It predicts the rate response to temperature, ferric and ferrous concentrations which are the important parameters during the second stage leaching.  136  CHAPTER 7 RECOMMENDATIONS FOR FURTHER WORK These studies have shed light on the conditions which favor the leaching o f chalcocite i n acidic ferric sulfate solution. The mineralogical studies have identified the formation o f intermediate copper sulfides and their probable roles during leaching have been discussed. Further mineralogical characterization o f the copper disulfide phase is desirable i n order to define this phase more accurately and to establish the conditions affecting its formation and stability. This may involve complete probing by E P M A , all the remaining grains after the leaching reaction rate rather than random selection o f the grains. The nature o f the elemental sulfur formed at different temperatures needs to be studied further i n order to establish its properties (physical and chemical) which may affect the second stage leaching. The focus o f this thesis was to carry out a fundamental investigation o f the effects of various leaching conditions on the kinetics o f chalcocite oxidation. A s a result, the fruitful combination o f leaching parameters to be employed so as to improve the operations o f heap and tank leaching operations may be inferred and the mathematical models for the leaching o f chalcocite are presented. However, the following suggestion is offered as a promising additional approach to pursue i n the future.  Use of Chloride The addition o f chloride ions to sulfate media i n copper heap leaching is being considered as one way to enhance copper dissolution. The dissolution reaction in chloride has been reported to be faster than that o f sulfate media. A l s o , it can be speculated that the chloride ion w i l l accelerate the decomposition o f the copper disulfide through the following reaction: CuS  2  + 2C1"  —>  C u C l " + 2 S ° +e" 2  (7-1)  There w i l l be need to establish the appropriate concentration o f total-chloride so as to prevent precipitation o f the C u C l " . 2  137  REFERENCES 1. Sullivan, J.D., Chemistry o f Leaching Covellite, U . S . Bureau o f Mines, Tech. Paper 487, 1930. 2. Marcantonio, P., Kinetics o f Dissolution o f Chalcocite i n Ferric Sulfate Solutions Ph.D. Thesis, Dept. o f M i n i n g , Metallurgical and Fuels Engineering, University o f Utah, Lake Utah, U . S . A , 1975. 3. M u l a k , W . , Kinetics o f Dissolving Poly dispersed Covellite i n A c i d Solutions o f Ferric Sulfate, Roczn.Chem., 45, pp. 1417-1424, 1971. 4.  L o w e , D . F . , The Kinetics o f the Dissolution Reaction o f Copper & Copper-Iron Sulfide Minerals Using Ferric Sulfate Solutions, Unpublished Ph.D. Thesis, University o f Arizona, U S A , 1970.  5. Tkachenko, O . B . and Tseft, A . L . , Kinetics o f the Dissolution o f Chalcocite i n Ferric Chloride, Trudy Inst. Metall. Obogashck, AlmaAta, 30, pp. 1387-1394, 1969. 6.  M u l a k , W . , Kinetics o f Cuprous Sulfide Dissolution i n A c i d i c Solutions o f Ferric Sulfate, Roczn. Chem., 43, pp. 1387-1394,1969.  7. K o p y l o v , G . A . and Orlov, A . 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W . , Comparison o f Chalcocite Dissolution i n the Sulfate, Perchlorate, Nitrate, Chloride, A m m o n i a and Cyanide Systems, Minerals Engineering, 7 (1), pp. 99-103, 1994. 74. Turney, T . . A . , Oxidation Mechanisms, Butterworth & C o (Publishers) L t d . , London, pp. 56, 90 and 111, 1965. 75. Tompkins, F . C . , The Kinetics o f the Reaction between Manganous and Permanganate ions, Transaction of the faraday Society, 38, pp. 131, 1942. 76. Jeffery, G . H . , Bassett, J., Mendham, J., and Denney, R . C , Vogel's Textbook of Quantitative Chemistry Analysis, Longman and John W i l e y , Newyork, N Y , U S A . 77. Silverman, M . P . and Lundgren, D . G . , Studies on the Chemoautotropic Iron Bacterium Ferrobacillus Ferrooxidans I. A n Improved M e d i u m and a Harvesting Procedure for Securing H i g h C e l l yields, Journal of Bacteriology, 77, pp. 642-647, 1959. 78. Burkin, A . R . , Composition and Phase Changes during Oxidative acid Leaching Reactions, Hydrometallurgical Process Fundamentals, R . G . Bautista, E d . , Plenum Press, N e w York, U S A , pp. 171-193, 1984. 79. Fogler, S. H . , Elements of Chemical Reaction Engineering, Second Edition, Chapter 10, External Diffusion Effects on Heterogeneous Reactions, Prentice H a l l International Series i n the Physical and Chemical Engineering Sciences, Prentice H a l l Upper Saddle River, N J . U S A , pp. 543-606, 1992. 80. Levenspiel, O., Chemical Reaction Engineering, Second Edition, Chapter 12, Fluidparticle Reactions, John Wiley & Sons, N e w York, N Y , U S A , pp. 357-408, 1972.  144  APPENDIX 1 Eh-pH DIAGRAM AT HIGH TEMPERATURE  In order to determine the AG °, the following are considered; 7  AG  298. A S ^  2 9 8  AG° = AH°-  (Al-1)  T.AS°  (Al-2)  and subtracting equation A l - 1 from equation A l - 2 , gives the following; AG". = A G  2 9 8  + (AH° - AH° ) - T.AS° + 29S.AS° T  29%  g&  (Al-3)  Recognising that, AHj — AH  (Al-4)  29&  298  =  AC^ .(T-29S) 9S  (Al-5)  In order to predict the heat capacity ( z l C / ^ ) , the following are considered; 98  Since  'dti\ C =, P  d  T  JdS_ j  (Al-6)  \~dT.  A t constant pressure; AS — Sj*  ^298  = 1A C -dr  (Al-7)  298  InT  =  (Al-8)  \AC° dInT P  7n298  V  Then,  AS° = AS° +  298  (Al-9) y  r  T  \  AC^l^M 298y  (Al-10)  V  B y substituting equation ( A l - 5 ) and ( A l - 1 0 ) into equation ( A l - 3 ) , then we obtain the following equation;  145  AG°  = AG°  T  GS  +  f  AC°  7 -298-77>i V  Where  Then,  e  T- 298 -Tin V  AG° = AG°  298  298  W  -(r-298)zLS °<298  —  9 S  2  (A-ll)  (A-12)  K29&J)  + AC° \ P  T  1  HGS  J9-(T-  298)AS  0 2 98  (A-13)  \c {T)dT p  IT  For the non-ionic species, CpJ  298  (A-14)  1  \dT 298  j(a + bT + cT~ yT 2  Therefore,  G ]1298 F  298  (A-15)  (7"-298)  a(r-298) + ^ ( r - 2 9 8 ) - c ( ^ - - ^ 2 1 T 298 2  2  V  ;  T-298  b{T+ 298) =a+•  c 298.7  146  (A-16)  (A-17)  The following are the free energies used to construct the E h - p H diagram for the C u - S - H 0 at 298 and 348 K ; 2  Table A l - 1 . Free energies (kJ/mol) o f the species used i n the construction o f the E h - p H diagrams. 298 K 348 K Specie Cu°  0.000  -1.211  Cu  2 +  65.700  69.665  Cu  +  50.300  47.633  HS0 " 4  -756.010  -761.680  S0 -  -744.600  -744.780  HS"  12.050  9.780  s-  86.310  87.930  Cu S ,  -87.600  -93.864  CuS  -53.200  -56.739  CuO  -134.000  -136.320  Cu 0  -148.100  -153.000  -33.560  -43.982  0.000  -1.526  2  4  2  2  2  H S (g) 2  S°  147  The following are the Heat capacity function used to construct the E h - p H at 348 K ;  Table A 1 - 2 . Heat capacity function o f the species (in J/mol-k) used i n the construction o f the E h - p H diagram. C A Specie b x 10 3  Cu°  22.635  6.276  0  Cu S  81.588  0  0  CuS  44.350  11.046  0  CuO  38.786  20.083  0  Cu 0  62.342  23.849  0  H S (g)  32.677  12.385  -1.925  S°  14.811  24.058  0.728  2  2  2  148  APPENDIX 2 EXPERIMENTAL DATA  Experimental Conditions: Variable  Temperature (chalcocite leaching)  Particle Size  -250+212 u m  Initial leach-volume  1000 m l  A c i d concentration  0.095 mol/L H S 0  Ferric concentration  0.116 mol/L  Ferrous concentration  0.020 mol/L  2  4  Initial mole o f C u i n the Solid mineral  0.0555 mol.  Table A 2 - 1 . Percentage o f copper extracted at different temperatures. Time (min)  35°C  55°C  65°C  75°C  0  0.00  0.00  0.00  0.00  1  9.10  14.69  20.81  34.01  3  20.23  25.47  31.79  40.21  4  N/A  29.00  35.00  41.52  5  28.00  32.25  36.82  42.53  6  N/A  35.08  38.85  43.75  8  N/A  37.77  40.95  44.15  10  34.70  39.68  42.04  46.46  15  37.77  41.88  44.09  48.23  149  Experimental Conditions: Variable  Ferric concentration (chalcocite leaching)  Temperature  35°C  Particle Size  -250+212 p m  Initial leach-volume  1000 m l  A c i d concentration  0.095 m o l / L H S 0  Ferrous concentration  0.010 mol/L  2  4  Initial mole o f C u in the Solid mineral  0.0555 mol.  Table A 2 - 2 . Percentage o f copper extracted at different concentrations o f ferric. Time (min) 0.007 M  0.015 M  0.029 M  0.058 M  0.116M  0.174 M  0.232 M  0  0.00  0.00  0.00  0.00  0.00  0.00  0.00  1  1.42  2.84  4.54  7.20  11.25  13.50  15.99  2  2.28  3.81  7.46  12.25  16.96  19.03  21.60  3  3.36  5.15  9.69  15.76  21.73  23.83  25.97  4  4.20  6.40  12.02  19.22  25.11  26.60  28.97  5  5.62  7.96  13.90  22.25  27.76  29.26  31.67  6  6.68  9.78  16.10  25.27  30.66  31.96  33.04  8  7.84  12.38  19.44  28.85  33.58  34.29  35.52  10  9.39  14.33  22.81  31.88  35.13  35.84  37.13  15  12.46  19.07  28.47  35.60  37.50  38.54  39.36  20  15.98  23.29  32.65  37.42  38.92  39.72  40.63  25  20.53  27.77  34.99  39.07  39.63  40.12  41.29  30  23.83  31.21  36.56  39.67  40.17  41.08  41.50  150  Experimental Conditions: Variable  Temperature (secondary-covellite leaching)  Particle Size  -250+212 u m  Initial leach-volume  1000 m l  A c i d concentration  0.095 mol/L H S 0  Ferric concentration  0.116 m o l / L  Ferrous concentration  0.0202 m o l / L  2  4  Initial mole o f C u in the Solid mineral  0.0555 mol.  Table A 2 - 3 . Percentage o f copper extracted at different temperatures. Time (hrs)  35°C  55°C  65°C  75°C  0.25  37.77  41.88  44.09  48.23  0.33  38.17  43.38  44.49  51.99  0.42  38.54  44.99  45.98  55.46  0.50  38.87  45.99  46.73  59.46  0.75  39.24  48.76  50.75  64.49  1.00  39.61  51.76  56.45  71.30  1.25  40.57  53.27  59.20  78.67  1.50  40.90  53.58  63.07  80.27  2.00  N/A  54.17  70.64  88.36  2.50  N/A  56.27  78.19  95.57  3.00  N/A  59.59  83.82  99.84  3.50  N/A  63.25  87.99  N/A  4.00  N/A  66.49  91.66  N/A  4.50  N/A  69.18  94.67  N/A  5.00  N/A  71.34  97.64  N/A  6.00  N/A  76.37  99.25  N/A  8.00  N/A  82.80  N/A  N/A  151  Experimental Conditions: Variable  Ferric (second stage leaching)  Particle Size  -250+212 p m  Initial leach-volume  1000 m l  A c i d concentration  0.095 m o l / L H S 0  Temperature  75°C  2  4  Ferrous concentration O.OlOlmol/L  Initial mole o f C u i n the Solid mineral  0.0555 m o l .  Table A 2 - 4 . Percentage o f copper extracted at different ferric levels. Time (hrs)  0.058 M  0.116M  0.232 M  0.348 M  0.05  29.43  36.97  38.07  38.77  0.08  37.76  39.55  40.53  41.85  0.13  39.95  43.48  44.82  45.39  0.25  46.30  48.75  50.09  51.37  0.33  49.86  51.62  53.44  55.94  0.42  51.62  54.19  56.80  58.51  0.50  54.11  57.86  60.04  61.61  0.75  61.24  64.17  67.89  69.39  1.00  67.71  71.69  74.24  75.94  1.25  74.71  78.79  81.30  83.06  1.50  80.65  84.84  87.98  89.25  2.00  86.71  91.70  94.75  96.23  2.50  90.56  N/A  N/A  N/A  3.00  93.61  N/A  N/A  N/A  .  152  Experimental Conditions: Variable  Ferrous (second stage leaching)  Particle Size  -250+212 u m  Initial leach-volume  1000 m l  A c i d concentration  0.095 m o l / L H S 0  Temperature  75°C  Ferric concentration  0.116 m o l / L  2  4  Initial mole o f C u i n the Solid mineral  0.0555 mol.  Table A 2 - 5 . Percentage o f copper extracted at different ferrous levels. Time (hrs)  0.001 M  0.010 M  0.054'M  0.101 M  0.25  51.85  53.63  49.96  40.55  0.33  55.36  56.78  N/A  N/A  0.42  59.74  59.61  N/A  N/A  0.50  61.48  62.49  58.90  55.42  0.75  69.62  69.30  66.36  N/A  1.00  77.47  77.42  N/A  68.55  1.25  82.35  83.51  78.56  N/A  1.50  89.17  89.93  83.57  79.89  2.00  98.82  99.03  93.68  89.30  2.25  N/A  N/A  98.32  N/A  2.50  N/A  N/A  N/A  98.12  153  Experimental Conditions: Variable  Ferric / ferrous ratio  Particle Size  -250+212 p m  Initial leach-volume  1000 m l  Total iron concentration  0.136 m o l / L  Temperature  75°C  Total sulfate concentration  0.287 m o l / L  Initial mole o f C u i n the Solid mineral  0.0555 m o l .  Table A 2 - 6 . Percentage o f copper extracted at different ferric / ferrous ratios. Time (hrs)  ferric  ferrous  Ferric  ferrous  ferric  Ferrous  0.116M  0.020 M  0.100 M  0.036 M  0.082 M  0.054 M  0.25  53.05  47.04  • i 9.90  0.33  57.19  N/A  N/A  0.42  61.01  N/A  N/A  0.50  65.41  57.81  54.85  0.75  69.46  64.76  63.38  1.00  78.43  76.26  68.29  1.25  86.54  82.16  76.95  1.50  92.09  88.74  79.80  2.00  99.53  97.58  90.33  2.50  N/A  99.39  93.98  /  154  APPENDIX 3 POTASSIUM PERMANGANATE ADDITION  The comparison between the extraction and the cumulative concentration o f potassium permanganate are illustrated in Figure A3-1 to Figure A 3 - 3 .  Figure A 3 - 1 . Plots o f copper extraction and K M n 0  4  concentration for the first stage  leaching o f -250+212 p m particles o f chalcocite at 35°C, 0.116 m o l / L ferric ion and 0.001 mol/L ferrous ion.  155  0  10  20  30  40  50  60  70  80  Time,hours  gure A 3 - 2 . Plots o f copper extraction and K M n 0 concentration for the second stage 4  leaching o f -250+212 \im particles o f chalcocite at 35°C, 0.116 m o l / L ferric ion and 0.001 mol/L ferrous ion.  156  100  0.014  gure A 3 - 3 . Plots o f copper extraction and K M n 0 concentration for the second stag 4  leaching o f -250+212 urn particles o f chalcocite at 75°C, 0.116 m o l / L ferric ion and 0.001 mol/L ferrous ion.  157  2.00-,  Time, hours  Figure A 3 - 4 . Effect o f K M n 0 on the leaching of -250+212 um particles of chalcocite 4  at 75°C, 0.000 m o l / L ferric, 0.095 m o l / L sulfuric acid and 0.015 m o l / L potassium permanganate.  158  APPENDIX 4 TOPOCHEMICAL MECHANISM OF LEACHING AND CALCULATION OF INSTANTANEOUS RATE OF L E A C H I N G  The  change  to the  surface  area as leaching progresses  is related to  the  instantaneous rate o f reaction and this is illustrated by plotting the rate against the percentage o f copper extracted as follows;  37.5 35.0 32.5"  T@55°C T@65°C T@75°C  30.0 27.5 25.0 22.5 -  i  20.0 -  £  ro  or  17.5 15.0 12.5 10.0 7.5 5.0 2.5 0.0 -  50  70  75  80  100  Extraction, %  Figure A 4 - 1 . Plots o f copper leach rate at different temperature vs. extraction for the second stage leaching o f -250+212 u m particles o f chalcocite at 75°C, 0.116 mol/L ferric ion and 0.0202 m o l / L ferrous ion.  159  In order to determine the instantaneous leach rates, the copper extraction curves (in Figure 4-31 were fitted to a polynomial function by using Microsoft  Excel  application. The polynomial function is o f the following general form;  y = Ax + A x + A  (A4-1)  2  {  2  3  where,  y = percent C u extraction, JC = leach residence time i n hours and A ,.A t  3  are constants.  The equation (A4-1) was differentiated to obtain the following rate equation;  dy -^ = 2A x + A ]  where,  (A4-2)  2  dy I dx is the instantaneous copper leach rate at any x,in units o f % C u / hour. Table A4-1 lists the values o f the coefficients obtained at different temperature.  Table A 4 - 1 . Coefficient values obtained for function fitted copper extraction data. A  A  Temperature (°C)  2  l  A  3  55  -0.2590  7.2177  41.774  65  -1.5807  19.7800  37.874  75  -5.7185  36.6990  39.967  160  

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