Open Collections

UBC Theses and Dissertations

UBC Theses Logo

UBC Theses and Dissertations

Atmospheric alkaline pre-oxidation of refractory sulphide gold ores Rego, Alan Joseph 2018

Your browser doesn't seem to have a PDF viewer, please download the PDF to view this item.

Item Metadata

Download

Media
24-ubc_2018_september_rego_alan.pdf [ 3.01MB ]
Metadata
JSON: 24-1.0366302.json
JSON-LD: 24-1.0366302-ld.json
RDF/XML (Pretty): 24-1.0366302-rdf.xml
RDF/JSON: 24-1.0366302-rdf.json
Turtle: 24-1.0366302-turtle.txt
N-Triples: 24-1.0366302-rdf-ntriples.txt
Original Record: 24-1.0366302-source.json
Full Text
24-1.0366302-fulltext.txt
Citation
24-1.0366302.ris

Full Text

Atmospheric Alkaline Pre-Oxidation of Refractory Sulphide Gold Ores  by  Alan Joseph Rego B.A.Sc., The University of British Columbia, 2012  A THESIS SUBMITTED IN PARTIAL FULFILLMENT OF THE REQUIREMENTS FOR THE DEGREE OF  MASTER OF APPLIED SCIENCE in THE FACULTY OF GRADUATE AND POSTDOCTORAL STUDIES (Materials Engineering)  THE UNIVERSITY OF BRITISH COLUMBIA (Vancouver)  May 2018  © Alan Joseph Rego, 2018ii  The following individuals certify that they have read, and recommend to the Faculty of Graduate and Postdoctoral Studies for acceptance, a thesis/dissertation entitled:  Atmospheric Alkaline Pre-Oxidation of Refractory Sulphide Gold Ores  submitted by Alan Joseph Rego  in partial fulfillment of the requirements for the degree of Master of Applied Science in Materials Engineering  Examining Committee: Dr. David B. Dreisinger, Materials Engineering Supervisor  Dr. Berend Wassink, Materials Engineering Supervisory Committee Member  Dr. David Dixon, Materials Engineering Supervisory Committee Member Dr. Warren Poole, Materials Engineering Additional Examiner   Additional Supervisory Committee Members:  Supervisory Committee Member  Supervisory Committee Member iii  Abstract  Four iron-sulphide-containing gold ores from the White Mountain mine in the Jilin Province, China were studied to examine the benefits of using oxidative treatment prior to conventional cyanidation to leach gold. This pre-treatment consisted of bubbling air into an ore/water mixture for eight hours while adding enough caustic sodium hydroxide to maintain a pH of 11 throughout the test. Three physical parameters were studied: total consumption of sodium hydroxide, rate of slurry aeration, and temperature of the slurry. Cyanide consumption was studied to determine whether increased oxidation of the ore would result in excessive degradation of sodium cyanide or otherwise impact the level of gold extraction. X-Ray Diffraction analysis was conducted on untreated samples as well as the final residues from the end of select tests to investigate by-products that precipitated after oxidation. Finally, an indicative economic analysis was conducted for each ore sample by comparing the increased dosage of caustic reagent applied during pre-treatment with the corresponding increased gold extraction. Results of this study show that aerated pre-oxidation of three of the four ores tested increased gold extraction compared to conventional cyanidation. Higher caustic dosages may have led to increased oxidation and were symptomatic of elevated release of sulphuric acid due to sulphide oxidation. Gold extraction was seen to rise in most cases after an increased rate of aeration during pre-treatment. Finally, the iron-sulphide ores tested responded extremely favourably to raised temperatures, showing both increased iron and sulphur oxidation, as well as higher gold extractions. Major losses of cyanide, likely due to production of cyanate and thiocyanate, were observed. Residual cyanide levels in the leach remained positive and did not affect gold extractions. Iron oxides such as hematite, and magnetite, and iron oxy-hydroxides such as goethite and ferrihydrite were observed in oxidized ore residues, indicating a high level of iron and sulphur oxidation.  iv  Lay Summary  Gold ores that are hard to process via conventional means are termed “refractory”. Often, there are undesired components in the ore that block physical contact between gold and leaching solution that extracts the gold. This work studies a method of removing problematic components in an ore that can occlude gold from leaching solution, thereby hindering gold extraction. Specifically, the method studied uses forced aeration in a slurry under high pH, ambient pressure conditions to oxidize iron-sulphide-containing gold ores taken from a mine in China. This work examines physical parameters that may improve gold extraction such as increased addition of caustic sodium hydroxide, different rates of air sparging in the slurry, and low to medium oxidation temperatures. Generally, all three of the above parameters were found to increase gold extractions in the ores, with the most improvement coming from higher temperature oxidation. v  Preface  This thesis is original, unpublished, and independent work by the author, Alan Joseph Rego.  All research reported herein was conducted by the author at the Point Grey (Vancouver, B.C.) campus of the University of British Columbia. Where noted, chemical analyses of samples were conducted by either Inspectorate International Ltd. (Bureau Veritas S. A.), or AuTec Innovative Extractive Solutions Ltd. (Barrick Gold Corporation) in Greater Vancouver, B.C.  This work was sponsored by Mitacs through the Mitacs-Accelerate Program and by the Eldorado Gold Corporation, which also provided the mineral ore samples for this work.      vi  Table of Contents  Abstract ................................................................................................................................................ iii Lay Summary ....................................................................................................................................... iv Preface ................................................................................................................................................... v Table of Contents ................................................................................................................................. vi List of Tables ......................................................................................................................................... x List of Figures ...................................................................................................................................... xii List of Symbols and Abbreviations .................................................................................................... xvi Acknowledgements ............................................................................................................................ xvii Dedication ......................................................................................................................................... xviii Chapter 1: Introduction ........................................................................................................................ 1 Chapter 2: Literature Review ............................................................................................................... 3 2.1 Physical Characteristics of Gold ..................................................................................................... 3 2.2 Chemistry of Gold Extraction ........................................................................................................ 4 2.2.1 Gold Leaching Using Conventional Methods........................................................................... 4 2.2.2 Gold Leaching Using Alternative Leaching Agents ............................................................... 12 2.2.2.1 Gold Leaching Using Thiosulphate ................................................................................. 13 2.2.2.2 Gold Leaching Using Thiourea ....................................................................................... 15 2.2.2.3 Gold Leaching Using Ammonia ..................................................................................... 16 2.2.2.4 Gold Leaching Using Halides ......................................................................................... 16 2.3 Processing of Refractory Gold Ores ............................................................................................. 17 2.3.1 Roasting of Iron Sulphide Minerals ....................................................................................... 18 2.3.2 Bio-oxidation of Iron Sulphide Minerals................................................................................ 19 vii  2.3.3 Aqueous Pressure Oxidation of Iron Sulphide Minerals ......................................................... 24 2.4 Physical and Chemical Characteristics of Iron Sulphide Minerals ................................................. 27 2.4.1 Pyrite and Marcasite .............................................................................................................. 27 2.4.2 Arsenopyrite ......................................................................................................................... 31 2.5 Atmospheric Oxidation of Iron Sulphide Minerals........................................................................ 36 2.5.1 Acidic Oxidation of Pyrite, Marcasite, and Arsenopyrite ....................................................... 37 2.5.2 Alkaline Oxidation of Pyrite, Marcasite, and Arsenopyrite .................................................... 42 Chapter 3: Experimental Work .......................................................................................................... 46 3.1 Materials...................................................................................................................................... 46 3.1.1 Reagents ............................................................................................................................... 46 3.1.2 Ore Samples .......................................................................................................................... 47 3.2 Preparation of Ore Samples .......................................................................................................... 51 3.3 Experimental Set-up .................................................................................................................... 52 3.4 Experimental Methods ................................................................................................................. 54 3.4.1 Oxidation Tests ..................................................................................................................... 55 3.4.2 pH Conditioning.................................................................................................................... 58 3.4.3 Cyanidation Tests .................................................................................................................. 58 Chapter 4: Results and Discussion ...................................................................................................... 60 4.1 Results of Baseline Cyanidation Tests .......................................................................................... 60 4.2 Results of Alkaline Oxidation Tests ............................................................................................. 64 4.2.1 Caustic Consumption ............................................................................................................ 66 4.2.1.1 Sample 165 .................................................................................................................... 66 4.2.1.2 Sample 167 .................................................................................................................... 69 4.2.1.3 Sample 169 .................................................................................................................... 71 4.2.1.4 Sample 201 .................................................................................................................... 73 4.2.2 Air Sparging ......................................................................................................................... 75 viii  4.2.2.1 Sample 165 .................................................................................................................... 76 4.2.2.2 Sample 167 .................................................................................................................... 77 4.2.2.3 Sample 169 .................................................................................................................... 79 4.2.2.4 Sample 201 .................................................................................................................... 82 4.2.3 Oxidation Temperature.......................................................................................................... 84 4.2.3.1 Sample 165 .................................................................................................................... 84 4.2.3.2 Sample 167 .................................................................................................................... 86 4.2.3.3 Sample 169 .................................................................................................................... 88 4.2.3.4 Sample 201 .................................................................................................................... 89 4.2.4 Cyanide Consumption ........................................................................................................... 91 4.2.4.1 Sample 165 .................................................................................................................... 92 4.2.4.2 Sample 167 .................................................................................................................... 93 4.2.4.3 Sample 169 .................................................................................................................... 94 4.2.4.4 Sample 201 .................................................................................................................... 95 4.3 Results of XRD Phase Analysis of Tails Residues ........................................................................ 97 4.3.1 Sample 165 ........................................................................................................................... 97 4.3.2 Sample 167 ........................................................................................................................... 98 4.3.3 Sample 169 ........................................................................................................................... 99 4.3.4 Sample 201 ......................................................................................................................... 100 4.4 Comparison of Ores and Discussion ........................................................................................... 101 4.5 Economic Analysis of Test Results ............................................................................................ 113 4.5.1 Sample 165 ......................................................................................................................... 115 4.5.2 Sample 167 ......................................................................................................................... 117 4.5.3 Sample 169 ......................................................................................................................... 119 4.5.4 Sample 201 ......................................................................................................................... 121 4.6 Reproducibility of Tests ............................................................................................................. 122 4.7 Summary of Results ................................................................................................................... 126 ix  Chapter 5: Conclusions and Recommendations ............................................................................... 129 5.1 Conclusions ............................................................................................................................... 129 5.2 Recommendations for Future Work............................................................................................ 132 References .......................................................................................................................................... 134 Appendices ......................................................................................................................................... 143 Appendix A: Analytical Methods Used to Test Leaching Samples ................................................... 143 A.1 Gold Fire Assays ................................................................................................................... 143 A.2 Sulphate Tests ....................................................................................................................... 143 A.3 Solid Sulphur Tests ............................................................................................................... 143 A.4 Aqueous Gold Tests .............................................................................................................. 144 A.5 Thiocyanate Tests ................................................................................................................. 144 Appendix B: XRD Imaging Results for Untreated Samples .............................................................. 145 B.1 Sample 165 ........................................................................................................................... 145 B.2 Sample 167 ........................................................................................................................... 146 B.3 Sample 169 ........................................................................................................................... 147 B.4 Sample 201 ........................................................................................................................... 148 Appendix C: Sample Calculations .................................................................................................... 149 C.1 Gold Extraction ..................................................................................................................... 149 C.2 Extraction Curves .................................................................................................................. 150 C.3 Calculated Head Values ......................................................................................................... 152 C.4 Cyanide Consumption ........................................................................................................... 152 C.5 Oxidation .............................................................................................................................. 154 Appendix D: Example Mass Balance Table ..................................................................................... 155 D.1 Sample 165 ........................................................................................................................... 156 Baseline Cyanidation Test #2 ................................................................................................... 156 High Temperature Oxidation (Test F):...................................................................................... 159 x  List of Tables  Table 1: Physical and Chemical Properties of Gold .................................................................................. 4 Table 2: Averaged Stability Data for Select Gold Complexes ................................................................. 13 Table 3: Physical and Chemical Properties of Pyrite .............................................................................. 28 Table 4: Physical and Chemical Properties of Marcasite ........................................................................ 29 Table 5: Physical and Chemical Properties of Arsenopyrite.................................................................... 32 Table 6: Chemical Reagents Used for Leaching Tests ............................................................................ 46 Table 7: Elemental Summary of All White Mountain Ore Samples ........................................................ 47 Table 8: Summary of Results from XRD Analysis and Rietveld Refinement (QXRD) ............................ 48 Table 9: Grind Calibration Data Obtained Through Sieving and Laser Size Analysis ............................. 51 Table 10: Conservative Grind Times Used to Obtain a P80 Size of 50 μm ............................................... 52 Table 11: Physical Parameters Used for Baseline Oxidation Tests .......................................................... 55 Table 12: Parameters Varied for Oxidation Tests ................................................................................... 56 Table 13: Physical Parameters Measured During Oxidation and Cyanidation Tests ................................ 56 Table 14: Physical Parameters Used for Cyanidation Tests .................................................................... 58 Table 15: Results of Baseline Cyanidation Tests on All Tested Ores ...................................................... 61 Table 16: Cyanide Consumption for Baseline Cyanidation Tests Conducted for All Ores ....................... 63 Table 17: Summary of Oxidation and Cyanidation Conditions Used for all Ore Samples........................ 65 Table 18: Total Cyanide Consumption and Breakdown for Sample 165 ................................................. 92 Table 19: Total Cyanide Consumption and Breakdown for Sample 167 ................................................. 93 Table 20: Total Cyanide Consumption and Breakdown for Sample 169 ................................................. 94 Table 21: Total Cyanide Consumption and Breakdown for Sample 201 ................................................. 95 Table 22: Effect of Measured Parameters on Total Cyanide Consumption .............................................. 96 Table 23: Effect of Measured Parameters on Thiocyanate Production .................................................... 96 Table 24: Results of XRD Analysis for Sample 165 (Excluding Clay Minerals) ..................................... 97 Table 25: Results of XRD Analysis for Sample 167 (Excluding Clay Minerals) ..................................... 98 Table 26: Results of XRD Analysis for Sample 169 (Excluding Clay Minerals) ..................................... 99 Table 27: Results of XRD Analysis for Sample 201 (Excluding Clay Minerals) ................................... 100 Table 28: Summary of Results for Alkaline Oxidation and Cyanidation ............................................... 103 Table 29: Summary of Economic Calculations for Sample 165 ............................................................ 115 Table 30: Summary of Economic Calculations for Sample 167 ............................................................ 117 Table 31: Summary of Economic Calculations for Sample 169 ............................................................ 119 Table 32: Summary of Economic Calculations for Sample 201 ............................................................ 121 xi  Table 33: Average Percent Difference Results for Recorded Data for Baseline Cyanidation Tests ........ 123 Table 34: Average Percent Difference Results for Recorded Data for Baseline Oxidation Tests ........... 123 Table 35: Percent and Absolute Difference of Final Extraction Values from Baseline Tests ................. 124 Table 36: Example Gold Calculations .................................................................................................. 150 Table 37: Example of Gold Concentrations with Dilution Factors ........................................................ 151 Table 38: Example of Extraction Curve Calculations ........................................................................... 151  xii  List of Figures  Figure 1: Historical Price Chart for Gold ($USD/kg and $USD/oz.) from Apr. 2015 – Apr. 2018 (Gold Price Group, 2017) .................................................................................................................................. 1 Figure 2: The atomic crystal structure of Gold (Crystallography365, 2014) .............................................. 3 Figure 3: Speciation diagram of 1M cyanide complexes in pure water at 25°C over a pH range of 0-14, by MEDUSA ............................................................................................................................................... 5 Figure 4: Pourbaix diagram for the thermodynamically stable CN-H2O system at 25°C with 1 mM cyanide concentration (Yannopoulos, 1991)............................................................................................. 6 Figure 5: Pourbaix diagram for the thermodynamically stable Au-H2O system at atmospheric pressure and 25°C temperature with 10-6 mM gold concentration by HSC 6.0 .............................................................. 7 Figure 6: Pourbaix diagram for the Au-CN-H2O system at atmospheric pressure and 25°C temperature with 1 mM gold concentration and various cyanide concentrations (Yannopoulos, 1991) ......................... 7 Figure 7: Pourbaix diagram for the thermodynamically stable S-H2O system at atmospheric conditions and 25°C with 0.1 M sulphur concentration by HSC 6.0 ........................................................................ 23 Figure 8: Pourbaix diagram for the thermodynamically metastable S-H2O system at atmospheric conditions and 25°C with 0.1 M sulphur concentration, and excluding some species such as sulphate, trithionate, and pentathionate, by HSC 6.0 ............................................................................................. 23 Figure 9: Face centered cubic crystal structure of pyrite; small black spheres indicate iron, larger grey spheres indicate sulphur (Uhlig, Szargan, Nesbitt, & Laajelehto, 2001) .................................................. 28 Figure 10: Face centered cubic crystal structure of marcasite; small black spheres indicate iron, larger grey spheres indicate sulphur (Uhlig, Szargan, Nesbitt, & Laajelehto, 2001) .......................................... 29 Figure 11: Pourbaix diagram for the thermodynamically stable Fe-H2O system at 25°C with 0.1 mM iron concentration by HSC 6.0 ...................................................................................................................... 30 Figure 12: Pourbaix diagram for the thermodynamically stable Fe-S-H2O system at 25°C with 0.1 mM for both iron and sulphur concentrations, by HSC 6.0 .................................................................................. 30 Figure 13: Pourbaix diagram for the thermodynamically stable As-H2O system at 25°C with 0.1 mM for arsenic concentration, by HSC 6.0 ......................................................................................................... 33 Figure 14: Pourbaix diagram for the thermodynamically stable As-S-H2O system at 25°C with 0.1 mM for both arsenic and sulphur concentrations, by HSC 6.0 ............................................................................. 33 Figure 15: Pourbaix diagram for the thermodynamically stable Fe-As-S-H2O system at 25°C with 0.1 mM for iron, arsenic, and sulphur concentrations, by HSC 6.0 ...................................................................... 34 Figure 16: Pourbaix diagram for the thermodynamically metastable Fe-As-S-H2O system at 25°C with 0.1 mM for iron, arsenic, and sulphur concentrations, omitting the scorodite phase, by HSC 6.0 .................. 34 xiii  Figure 17: Complex Pourbaix diagram for the thermodynamically stable Fe-As-S-H2O system at 25°C with 0.1 mM for iron, arsenic, and sulphur concentrations (Bhakta, Langhans, & Lei, 1989). ................. 35 Figure 18: XRD Pattern for untreated ("Head") sample 165; analysis and Rietveld Refinement performed by the department of Earth and Ocean Sciences at UBC ......................................................................... 49 Figure 19: Dried ore samples after crushing and grinding....................................................................... 50 Figure 20: Plot of measured P80 values obtained using the sieving and Laser Size Analysis techniques ... 52 Figure 21: Schematic of the leaching reactor used, with lid attached, and mixing unit suspended above. The holes drilled into the lid were ordered as follows: 1) Air sparging inlet; 2) Dissolved Oxygen electrode; 3) Air vent (if needed); 4, 5, 6) Eh-pH probes and sampling outlet; 7) Impeller shaft inlet ...... 53 Figure 22: Schematic of the layout of the experimental set-up (overhead view). Components are numbered: 1) Reactor; 2) Heating element; 3) Thermometer; 4) Overhead mixer unit; 5) Humidifier ..... 53 Figure 23: Schematic of a sealed Erlenmeyer flask, filled with De-Ionized (DI) water, used as a humidifier. Components are numbered as follows: 1) Air input hose; 2) Air dispersion tube and glass frit; 3) Humidified air outlet hose ................................................................................................................. 54 Figure 24: Baseline cyanidation curves for all tested ores; from left to right: sample 165, 167, 169, 169; (35°C temperature, 20% solids by mass (initial condition), no air sparging, pH > 11) ............................. 61 Figure 25: Extent of oxidation vs. caustic consumption for sample 165 .................................................. 66 Figure 26: Gold extraction vs. extent of oxidation for sample 165 .......................................................... 67 Figure 27: Gold extraction vs. caustic consumption for sample 165; alkaline oxidation tests results only 68 Figure 28: Gold extraction vs. caustic consumption for sample 165; baseline cyanidation results included.............................................................................................................................................................. 68 Figure 29: Extent of oxidation vs. caustic consumption for sample 167 .................................................. 69 Figure 30: Gold extraction vs. extent of oxidation for sample 167; baseline cyanidation data lie at 0% oxidation ............................................................................................................................................... 70 Figure 31: Gold extraction vs. caustic consumption for sample 167; baseline cyanidation results included.............................................................................................................................................................. 71 Figure 32: Extent of oxidation vs. caustic consumption for sample 169 .................................................. 72 Figure 33: Gold extraction vs. extent of oxidation for sample 169; baseline cyanidation results included 72 Figure 34: Gold extraction vs. caustic consumption for sample 169; baseline cyanidation results included.............................................................................................................................................................. 73 Figure 35: Extent of oxidation vs. caustic consumption for sample 201 .................................................. 74 Figure 36: Gold extraction vs. extent of oxidation for sample 201; baseline cyanidation assumed to have no oxidation .......................................................................................................................................... 74 xiv  Figure 37: Gold extraction vs. caustic consumption for sample 201; baseline cyanidation results included.............................................................................................................................................................. 75 Figure 38: Extent of oxidation vs. rate of air input at 0.68 atm (10 psi) for sample 165 ........................... 76 Figure 39: Gold extraction vs. rate of air input at 0.68 atm (10 psi) pressure for sample 165; baseline cyanidation results included ................................................................................................................... 77 Figure 40: Extent of oxidation vs. rate of air input at 0.68 atm (10 psi) for sample 167 ........................... 78 Figure 41: Gold extraction vs. rate of air input at 0.68 atm (10 psi) pressure for sample 167; baseline cyanidation results included ................................................................................................................... 78 Figure 42: Extent of oxidation vs. average dissolved oxygen in solution during oxidation for sample 167.............................................................................................................................................................. 79 Figure 43: Extent of oxidation vs. rate of air input at 0.68 atm (10 psi) for sample 169 ........................... 80 Figure 44: Gold extraction vs. rate of air input at 0.68 atm (10 psi) for sample 169; baseline cyanidation results included ..................................................................................................................................... 81 Figure 45: Extent of oxidation vs. average dissolved oxygen in solution during oxidation for sample 169.............................................................................................................................................................. 81 Figure 46: Extent of oxidation vs. rate of air input at 0.68 atm (10 psi) pressure for sample 201 ............. 82 Figure 47: Gold extraction vs. rate of air input at 0.68 atm (10 psi) pressure for sample 201; baseline cyanidation results included ................................................................................................................... 83 Figure 48: Extent of oxidation vs. average dissolved oxygen in solution during oxidation for sample 201.............................................................................................................................................................. 83 Figure 49: Extent of oxidation vs. oxidation temperature for sample 165................................................ 85 Figure 50: Gold extraction vs. oxidation temperature for sample 165 ..................................................... 86 Figure 51: Extent of oxidation vs. oxidation temperature for sample 167................................................ 87 Figure 52: Gold extraction vs. oxidation temperature for sample 167 ..................................................... 87 Figure 53: Extent of oxidation vs. oxidation temperature for sample 169................................................ 88 Figure 54: Gold extraction vs. oxidation temperature for sample 169 ..................................................... 89 Figure 55: Extent of oxidation vs. oxidation temperature for sample 201................................................ 90 Figure 56: Gold extraction vs. oxidation temperature for sample 201 ..................................................... 90 Figure 57: Gold extraction vs. extent of oxidation for tested ores; baseline cyanidation results included108 Figure 58: Extent of oxidation vs. total caustic consumed for tested ores; baseline cyanidation results included .............................................................................................................................................. 109 Figure 59: Gold extraction vs. total caustic consumed for tested ores; baseline cyanidation results included (plotted on x=0 for sample 165, shown as the data with the lowest caustic dosages for ores 167 and 169, and at 14.6 kg/t and 24.4 kg/t caustic dosage for sample 201, respectively) ............................ 109 xv  Figure 60: Extent of oxidation vs. rate of air input at 0.68 atm (10 psi) pressure for tested ores; alkaline oxidation results only .......................................................................................................................... 110 Figure 61: Gold extraction vs. rate of air input at 0.68 atm (10 psi) pressure for tested ores; baseline cyanidation results included ................................................................................................................. 110 Figure 62: Extent of oxidation vs. dissolved oxygen in solution during oxidation for tested ores; alkaline oxidation results only; trendline plotted for all ores except sample 165 ................................................ 111 Figure 63: Extent of oxidation vs. oxidation temperature for tested ores; alkaline oxidation results only............................................................................................................................................................ 112 Figure 64: Gold extraction vs. oxidation temperature for tested ores; alkaline oxidation results only .... 112 Figure 65: Extent of oxidation vs. average electrochemical potential observed during oxidation for tested ores; alkaline oxidation results only; trendline plotted for all ores except sample 165 ........................... 113 Figure 66: Potential Achievable Net Revenue Per Tonne Ore for All Ores Tested ................................ 114 Figure 67: XRD Pattern for untreated ("Head") sample 165; analysis and Rietveld Refinement performed by the department of Earth and Ocean Sciences at UBC ....................................................................... 145 Figure 68: XRD pattern for untreated ("Head") sample 167; analysis and Rietveld Refinement performed by the department of Earth and Ocean Sciences at UBC ....................................................................... 146 Figure 69: XRD pattern for untreated ("Head") sample 169; analysis and Rietveld Refinement performed by the department of Earth and Ocean Sciences at UBC ....................................................................... 147 Figure 70: XRD pattern for untreated ("Head") sample 201; analysis and Rietveld Refinement performed by the department of Earth and Ocean Sciences at UBC ....................................................................... 148 Figure 71: Example gold extraction curve ............................................................................................ 152  xvi  List of Symbols and Abbreviations  AAS    Atomic Adsorption Spectroscopy ACS    American Chemical Society atm    Pressure, in standard atmospheres (or 101,325 Pa) Ci   Concentration of element or compound in solution at time i DI    De-Ionized water DO    Dissolved Oxygen in solution| e-    Electron  Eh    Electrochemical potential, in Volts (V) E°    Standard electrochemical potential at 25°C, in Volts (V) ΔGf°298K   Standard Gibbs Free Energy (of formation at 298K or 25°C), in kJ/mol ICP-AES/OES  Inductively Coupled Plasma Atomic Emission or Optical Emission Spectroscopy ICP-MS  Inductively Coupled Plasma Mass Spectrometry LECO   Combustion analysis of total sulphur content using LECO® instruments LPM    Litres Per Minute M    Molarity (mol/L) mol   Moles of element or compound mt    Metric Tonne (1000 kg) N    Normality (eq/L) ORP    Oxidation/Reduction Potential, in Volts (V) pH    Measure of acidity/alkalinity in solution ppm    Parts Per Million, also equivalent to grams per tonne psi    Pressure, in pounds per square inch (lbs./in.2) PLS   Pregnant Leach Solution, the aqueous leachate solution at the end of leaching P80    the diameter of screen hole that allows 80% of ore particles to pass through RPM    Rotations Per Minute SEM    Scanning Electron Microscopy SHE    Standard Hydrogen Electrode Vi   Volume of solution aliquot or sample at time i XRD    X-Ray Diffraction analysis β    Stability constant of aqueous species  xvii  Acknowledgements  My most heartfelt thanks and gratitude go to my supervisor, Dr. David Dreisinger, for the support, advice, knowledge, and experience he has shared with me to aid in the completion of this work. Most of all, I am grateful for his near-infinite patience with me as I struggled to the finish line.  To Dr. Berend Wassink, my sincere gratitude for the practical advice and assistance he has provided in helping me set up experiments, analyze data, work through problems, and write this thesis. I am grateful for the safety training and dangerous materials handling suggestions that were invaluable to me throughout this undertaking.  Thanks go to Jacob Kabel for his assistance and advice pertaining to imaging my samples and troubleshooting the various issues that I faced in this endeavour. Thank you to Dr. Pius Lo, Jophat Engwayu, and Aaron Hope, of the Mining Engineering department, for helping me go through the physical processing of my samples, and for ensuring my safety throughout those long days.  Thank you to all the Faculty and Staff of the Materials Engineering department that helped me along throughout my journey – Fiona Webster, Mary Jansepar, Norma Donald, Sherry Legislador, Michelle Tierney, Marlon Blom, Ross McLeod, Carl Ng, David Torok, Wongsang Kim, Dr. Jianming Lu, and all those who I may have missed here.  Thank you to Dr. Jinxing Ji, Hai Guo, and the Eldorado Gold Corporation for not only the financial assistance necessary for this project, but also the critical information and feedback given at the onset of this work. My sincere thanks also go to the Mitacs Accelerate program and their representative, Sherry Zhao, for the matching financial contribution for this work, and extracurricular training provided.  To my friends and colleagues, both in the department and those outside, your friendship has meant the world to me. Thank you for your compassion and understanding, and for walking with me through this experience. xviii  Dedication  I dedicate this work to my Lord and my God. Thank you, Father, for forming me, for saving me, and for sanctifying me. Thank you for all your love and blessings, for all your mercy and kindness. Teach me to love you more and more each day.  “I can do all things through him who strengthens me.”  – Philippians 4:13  I also dedicate this work to my family. To Mum, Dad, Mamana, Molly, and especially Rachel: I love you. Without you I am nothing. Thank you for pushing me to be the best version of myself, and for holding me accountable when I need it the most. Thank you for your prayers, for your love, for your support, for listening to me complain, and for proof-reading all my boring reports; for everything that you have done for me throughout the last twenty-eight years of my life, and for all that is to come.  “… The truth is obtained like gold, not by letting it grow bigger, but by washing off from it everything that isn’t gold.” – Lev Nikolayevich Tolstoy    “For where your treasure is, there your heart will be also.”  – Luke 12:34 1  Chapter 1: Introduction  Gold is one of the world’s rarest and most valuable commodities. Achieving a record price of over $1900 USD/oz. in 2011, gold bullion has had an average price of approximately $1200 USD/oz. for at least the past three years (Gold Price Group, 2017). Given below in Figure 1 is the price history chart for a 3-year timescale.  Figure 1: Historical Price Chart for Gold ($USD/kg and $USD/oz.) from Apr. 2015 – Apr. 2018 (Gold Price Group, 2017) Gold resources (ore deposits) have become increasingly rare due to extensive efforts to extract the precious metal, and the quality of newly discovered deposits are becoming leaner and harder to process by conventional methods (Klein, 2010; Marsden & House, 2006). The gold mining industry is constantly looking to find new ways of extracting precious metals, and increasing their recoveries through novel metallurgical techniques (Farley, 1998; La Brooy, Linge, & Walker, 1994; Marsden & House, 2006; Prasad, Mensah-Biney, & Pizarro, 1991; Yannopoulos, 1991). In one such example, the White Mountain mine, formerly owned by the Eldorado Gold Corporation, used a high pH (alkaline) pre-treatment process to increase gold extraction from sulphide ores mined from the Jilin Province, located in North-East China. Sodium hydroxide (NaOH) addition and sparging with air into ore slurries effected the oxidation of sulphide minerals, resulting in a higher extraction of gold at the end of the process. While the overall effect was positive, this process was not clearly understood, and thus required further study to describe the various factors that may influence the oxidation of sulphide ores in alkaline media and the resulting effects on gold extraction. 2  Thus, this work was undertaken with the following goals: 1. To record rates of oxidation of ore samples obtained from the White Mountain mine and relating this back to the varying contents of pyrite, marcasite, arsenopyrite, silicate material, as well as any other significant gold-containing minerals in each ore. Specifically: a. Measure the rates of reaction and develop models to confirm the mode of oxidation b. Determine the controlling mechanism of the reaction and distinguish between factors that control the rate of reaction (e.g. pyrite and marcasite oxidation conforming to various leaching models such as mass-transfer through ion-diffusion, or chemical reaction control) c. Determine the activation energies for iron sulphide minerals in each sample of ore that is tested 2. To delineate the effects of alkaline oxidative pre-treatment on the extraction of gold from each natural ore sample by cyanidation. Parameters investigated include caustic dosage, air sparging levels, and temperature. 3. To describe the oxidation products that resulted from the alkaline pre-treatment step, and any influence they may have on the overall extraction of gold To complete this research, leaching studies were performed on four selected ore samples obtained from the White Mountain mine. Experiments consisted of alkaline pre-treatment, and cyanidation at slightly elevated temperature. In some cases, X-Ray diffraction was then undertaken to analyze and characterize the residue remaining from the leaching and cyanidation tests. This thesis is comprised of a literature review, found in Chapter 2, an explanation of all experimental work, set out in Chapter 3, experimental results and relevant discussion in Chapter 4, and finally, conclusions and recommendations in Chapter 5.3  Chapter 2: Literature Review  Given below is some relevant discussion of the various aspects of gold leaching, processing of refractory gold ores, including those with iron sulphide minerals, and atmospheric oxidation of iron sulphide minerals prior to cyanidation.  2.1 Physical Characteristics of Gold  Gold is a ductile and malleable metal, resistant to nearly all forms of corrosion. Gold derives its value from having an extremely pleasant lustre and appearance. While gold is rarely oxidized in ambient natural conditions, it is sometimes associated in nature with silver, forming a gold-silver alloy called “Electrum”, which was formed during geologically active times when intense heat and pressure forces melted gold and silver together (La Brooy, Linge, & Walker, 1994). Native gold has a crystal structure known as “face-centred cubic”, which can be seen simulated in Figure 2 below (Crystallography365, 2014):  Figure 2: The atomic crystal structure of Gold (Crystallography365, 2014) Gold is difficult to oxidize under natural atmospheric conditions. Gold is therefore considered to be extremely “noble”, and this fact is reflected by the high electrochemical reduction potentials (E°, under standard conditions, or Eh) that is required to drive electron transfer reactions of gold to completion. A compiled summary of the physical and chemical properties of gold is given below in Table 1 (The Chemical Rubber Company, 2014; Jara & Bustos, 1992; Marsden & House, 2006):  4  Table 1: Physical and Chemical Properties of Gold Property Value Unit Atomic Weight 196.97 g/mol Crystal Structure Face-Centered Cubic (FCC)  Density 19300 kg/m3 Melting Point 1064.18 (1337.33) °C (K) Heat Capacity 25.418 J/(mol∙K) Electrical Resistivity (20°C) 22.14∙10-3 Ω∙m Thermal Conductivity 318 W/(m∙K) Mohs Hardness 2.5  E° (Au+3) 1.5 (1.52, from Jara & Bustos, 1992)) V SHE E° (Au+) 1.69 (1.83, from Jara & Bustos, 1992) V SHE E° (Au(CN)2-) -0.6 V SHE  2.2 Chemistry of Gold Extraction 2.2.1 Gold Leaching Using Conventional Methods  Gold is typically produced via the MacArthur-Forrest technique, also known as “Cyanidation” (Klein, 2010; Logsdon, Hagelstein, & Mudder, 1999; Marsden & House, 2006) This is currently the most commonly used procedure around the world, with over 90% of all gold being extracted from mineral ore in this way (Farley, 1998; International Cyanide Management Institute, 2012; Klein, 2010; Laitos, 2012; Logsdon, Hagelstein, & Mudder, 1999; Marsden & House, 2006; Prasad, Mensah-Biney, & Pizarro, 1991). The cyanidation process uses a cyanide (CN-) salt to act as a ligand whereby gold is complexed with cyanide in aqueous systems (Klein, 2010; Kondos, Deschenes, & Morrison, 1995; Marsden & House, 2006) Typically, this cyanide compound is sodium cyanide, however other cations such as potassium may also be found in use with the gold-mining industry, depending on the price and availability of such commercially used reagents. Before discussing the chemistry of gold cyanidation, it is important to specify the conditions under which cyanidation typically occurs. Cyanide is a highly toxic compound composed of one carbon atom triple bonded to a nitrogen atom (International Union of Pure and Applied Chemistry, 2014). Cyanide ions will form hydrogen cyanide 5  (HCN) gas at sufficiently low pH, according to the following reaction (Logsdon, Hagelstein, & Mudder, 1999; Young & Jordan, 1995): CN-(aq) + H+(aq) = HCN(aq)        (eq. 1) Hydrogen cyanide is an extremely deadly gas, which has been known to kill humans within minutes of exposure, depending on the amount of gas inhaled, and the duration of exposure (International Cyanide Management Institute, 2012).  To prevent the formation of hydrogen cyanide, the use of cyanidation must occur in alkaline conditions. In practice, cyanidation takes place above a pH of approximately 10.5, giving practitioners of the method a safety margin of one unit of pH. This is because the level of formation of hydrogen cyanide gas from aqueous cyanide is 50% at a pH of approximately 9.31. Below a pH of 9.31, more than 50% of aqueous cyanide will tend to form HCN and at a pH of 8.2, over 90% of cyanide initially found in solution may form HCN gas. Figure 3 and Figure 4 below illustrates this concept more clearly (Puigdomenech, 2010; Yannopoulos, 1991).  For this work, all Eh-pH diagrams were drawn using the HSC Chemistry 6.0 software by Outokompu Research Oy, unless otherwise stated (Outokompu Research Oy, 2006). All speciation diagrams were drawn using the MEDUSA software by KTH Royal Institute of Technology, unless otherwise stated (Puigdomenech, 2010). Unless stated otherwise, all Eh values are given in Volts, relative to the Standard Hydrogen Electrode (SHE).  Figure 3: Speciation diagram of 1M cyanide complexes in pure water at 25°C over a pH range of 0-14, by MEDUSA 6   Figure 4: Pourbaix diagram for the thermodynamically stable CN-H2O system at 25°C with 1 mM cyanide concentration (Yannopoulos, 1991) Note that in the Pourbaix diagram below, both HCN and the cyanide ion (CN-) are shown to be chemically unstable at a pH > 4 at natural Eh (0V SHE), and that the cyanate (OCN-(aq)) species has a wide area of stability. In practice, aqueous cyanide is very kinetically stable, only being oxidized to cyanate after considerable time has passed under relatively high electrochemical potentials (Eh). The Pourbaix diagram for the Au-H2O system at atmospheric pressure and standard temperature (25°C) is given here in Figure 5. For Figures 5 and 6 below, the concentration of aqueous gold is at 10-6 M, a very low concentration given that gold is typically found in natural ores at ppm (g/mt) levels. Assuming a very low value of 1g Au per metric tonne and considering the test parameters used for this research (100g ore, 0.5L leaching solution) this gives a minimum molarity of 10-6M Au. Figure 6 gives much higher gold concentrations as well for comparison. 7   Figure 5: Pourbaix diagram for the thermodynamically stable Au-H2O system at atmospheric pressure and 25°C temperature with 10-6 mM gold concentration by HSC 6.0 Given below in Figure 6 is the Pourbaix diagram for the Au-CN-H2O system (Yannopoulos, 1991).  Figure 6: Pourbaix diagram for the Au-CN-H2O system at atmospheric pressure and 25°C temperature with 1 mM gold concentration and various cyanide concentrations (Yannopoulos, 1991) 8  As shown above, a large area is dominated by the presence of the dicyanoaurate (I) complex [Au(CN)2-]. The presence of the CN- ligand in an alkaline system enables gold to easily form a stable aqueous complex that is then amenable to further processing. This means that gold can be complexed from a solid state in mineral ores to an aqueous state under the right conditions. Subsequently, the aqueous gold can be concentrated and reconstituted into a solid product for sale or refinement.  The MacArthur-Forrest process (U.S.A. Patent No. 403202, 1889; U.K. Patent No. 14174, 1887) by which gold is dissociated into an aqueous solution has been the focus of intense study for decades, and rightly so, for the dividends that may be gained from extracting one percentage point more gold from an ore deposit may sum upwards of hundreds of thousands of dollars. The overall chemistry is given by the Elsner equation (Marsden & House, 2006), written below: 4Au(s) + 8NaCN(aq) + O2(aq) + 2H2O(l) = 4Na[Au(CN)2](aq) + 4NaOH(aq)   (eq. 2) The Elsner equation is a combination of a reduction reaction and an oxidation reaction. This combination, known as a “redox” reaction, is separated into its two components below (Multi Mix Systems Pty. Ltd., 2009; Senanayake, 2005; Wadsworth M. E., 2000): Reduction: Step 1: O2(aq) + 2H2O(l) + 2e- = H2O2(aq) + 2OH-(aq)     (eq. 3a) Step 2: H2O2(aq) + 2e- = 2OH-(aq)        (eq. 3b) Overall Reduction: O2(aq) + 2H2O(l) + 4e- = 4OH-(aq)  E° = 0.401 V SHE (eq. 3) Oxidation: Step 1: Au(s) = Au+(aq) + e-        (eq. 4a) Step 2a: Au+(aq) + CN-(aq) = Au(CN)(aq)       (eq. 4b) Step 2b: Au(CN) + CN-(aq) = Au(CN)2-(aq)      (eq. 4c) Overall Oxidation: Au(s) + 2CN-(aq) = Au(CN)2-(aq) + e-  E° = -0.57 V SHE (eq. 4) As shown above, the anodic dissolution occurs around -0.57 V SHE (Guzman, Segarra, Chimenos, Cabot, & Espiell, 1999).  To balance the reactions when both are added to each other, the overall reduction reaction is added to four times the oxidation reaction: Overall Redox = (1)(Reduction) + (4)(Oxidation) 4Au(s) + 8CN-(aq) + O2(g) + 2H2O(l) = 4Au(CN)2-(aq) + 4OH-(aq) E° = -0.169 V SHE (eq. 2a) 9  Note that the above reactions delineate gold oxidation as a separate sub-step (i.e. Au is oxidized to Au(I)). Many authors propose a different mechanism, whereby gold is first complexed with cyanide to become an adsorbed species, which either simultaneously or subsequently releases an electron as the oxidation half reaction (Nicol M. J., The Anodic Behaviour of Gold Part II - Oxidation in Alkaline Solutions, 1980; Senanayake, Kinetics and Reaction Mechanism of Gold Cyanidation: Surface Reaction Model via Au(I)-OH-CN Complexes, 2005; Wadsworth, Zhu, Thompson, & Pereira, 2000). These mechanisms contend that an intermediate “metastable” complex is formed, either Au(CN)-(ads), or Au(CN)(ads), depending on when oxidation of the gold species occurs in the leaching process. With these physiochemical mechanisms, cyanide works to first adsorb onto the surface of gold (rather than first oxidizing gold to a higher valence), lowering the electrochemical potential needed to extract gold from the bulk precious metal grain. The oxidation of gold in an alkaline cyanide solution has been studied extensively, with numerous electrochemical, physio-chemical, and surface reaction mechanisms proposed by various authors. For the reaction mechanism shown above, the oxidation of gold occurs in 2 main stages: the actual oxidation of gold to an aurous (Au+) species, and then the complexation of this aurous species with cyanide in a step-wise fashion, until it detaches from the surface of the bulk gold particle as the dicyanoaurate (I) complex (Kondos, Deschenes, & Morrison, 1995; Marsden & House, 2006; Multi Mix Systems Pty. Ltd., 2009; Senanayake, Kinetics and Reaction Mechanism of Gold Cyanidation: Surface Reaction Model via Au(I)-OH-CN Complexes, 2005; Wadsworth, Zhu, Thompson, & Pereira, 2000). Elsner showed that oxygen is critical for the cyanidation reaction to proceed (Jara & Bustos, 1992; Marsden & House, 2006; Senanayake, “Kinetics…”, 2005; Wadsworth M. E., 2000; Zhang, Fang, & Muhammed, 1997). It is theorized that oxygen acts as an oxidant, inducing native gold (Au) to release either one electron, or three electrons, depending on the electrochemical potential of the system. The reduction half-reactions are given below: Au+(aq) + e- = Au(s)    E° = 1.83 V (a), 1.691 V (b)  (eq. 5) Au+3(aq) + 3e- = Au(s)    E° = 1.52 V (a), 1.498 V (b)  (eq. 6) a. (Jara & Bustos, 1992) b. (Marsden & House, 2006; Nicol M. J., The Anodic Behaviour of Gold Part II - Oxidation in Alkaline Solutions, 1980)  The kinetics of gold extraction using cyanide, or indeed any other leaching reagent, is most likely dictated by the limitations of mass transfer, meaning that the rate-limiting step of gold leaching is probably the diffusion of these leaching agents to the surface of gold particles (or “grains”) (Deschenes, 2005; Guzman, 10  Segarra, Chimenos, Cabot, & Espiell, 1999; Klein, 2010; Marsden & House, 2006). However, there is evidence that surface electrochemical reactions control the cyanidation process under certain conditions (Guzman, Segarra, Chimenos, Cabot, & Espiell, 1999; Senanayake, Kinetics and Reaction Mechanism of Gold Cyanidation: Surface Reaction Model via Au(I)-OH-CN Complexes, 2005; Wadsworth, Zhu, Thompson, & Pereira, 2000; Wadsworth M. E., 2000). This evidence arises from discrepancies between activation energies reported for the dissolution of gold in a cyanide solution (47 – 55 kJ/mol) and normal activation energies associated with diffusion controlled reactions (5 – 20 kJ/mol) (Wadsworth, Zhu, Thompson, & Pereira, 2000) as well as lower reported dissolution rates at varying test conditions, indicating that, even in solutions with greater cyanide concentrations and oxygen saturation levels, other rate-limiting factors may play a role in the overall cyanidation mechanism (Senanayake, “Kinetics…”, 2005; Wadsworth M. E., 2000). At low cyanide concentrations, cyanidation is controlled by diffusion of cyanide molecules (CN-(aq)) from the bulk solution to the surface of a gold particle; however, at higher cyanide concentrations, the diffusion of oxygen from bulk solution to the surface of gold particles controls the rate of reaction (Jara & Bustos, 1992; Kondos, Deschenes, & Morrison, 1995; Senanayake, Kinetics and Reaction Mechanism of Gold Cyanidation: Surface Reaction Model via Au(I)-OH-CN Complexes, 2005). Special consideration must be given to saturate slurries with oxygen to ensure maximum concentration of aqueous oxygen in the system. This is because the amount of dissolved oxygen typically found in solution during cyanidation (or any oxidative leaching process) is quite low. Because oxygen saturation is limited, the reactions that require reduction of water are slower due to the limitation of oxygen diffusing to the surface of the reacting metal. Additionally, efforts to saturate solutions with oxygen through forced aeration (bubbling of air) are not very efficient: where air is being bubbled through an aqueous system, 10 – 20% of all oxygen pumped into the system may be retained in solution as aqueous or Dissolved Oxygen (D.O.), and this range can be further influenced by the various parameters that affect oxygen solubility in a solution (such as temperature, salt content, etc.) In dilute cyanide liquors, a maximum oxygen level of 8.2 ppm (or mg/L) is achievable (Deschenes, 2005), however this is with manual aeration (air pumped into solution). Conventional cyanidation occurs with a dissolved oxygen (D.O.) content of approximately 6 ppm (Deschenes, 2005). While the presence of dissolved oxygen in a leaching solution is crucial to the completion of the gold extraction process through the Elsner reaction, there is a downside to excess oxygenation. Increased oxygen in solution may lead to an increased breakdown of cyanide to cyanate (CNO-) under alkaline conditions; this is shown by the reactions below (Marsden & House, 2006; Young & Jordan, 1995): Breakdown of cyanide to cyanate by reaction with hydroxide Oxidation: CN-(aq) + 2OH-(aq) = CNO-(aq) + H2O(l) + 2e-     (eq. 7a) 11  Reduction: O2(aq) + H2O(l) + 4e- = 4OH-(aq)      (eq. 7b) Overall Redox = (2)(a) + (1)(b)  2CN-(aq) + O2(aq) = 2CNO-(aq)        (eq. 7) Cyanide can also be consumed via reaction with native sulphur (S8) to produce thiocyanate (SCN-) in addition to other aqueous species such as thiosulphate (S2O3-2), sulphite (SO3-2), depending on the extent of oxidation of sulphur (Jones, 1998). Reactions between various sulphur compounds are shown below (Jones, 1998; Miller & Brown, 2005; Yannopoulos, 1991; Young & Jordan, 1995): Production of thiocyanate by reaction of cyanide with native sulphur CN-(aq) + S80(s) = SCN-(aq) + S70(s)        (eq. 8a) Overall Reaction: 8CN-(aq) + S8(s) = 8SCN-(aq)      (eq. 8) Production of thiocyanate by reaction of cyanide with thiosulphate CN-(aq) + S2O3-2(aq) = SCN-(aq) + SO3-2(aq)       (eq. 9) Or: 2CN-(aq) + 2S2O3-2(aq) + O2(aq) = 2SCN-(aq) + 2SO4-2(aq)     (eq. 10) The latter reaction described is a combination of two separate reactions: the formation of thiocyanate from sulphur and cyanide, and the destruction of thiosulphate forming sulphate.   Oxidation of thiocyanate to sulphate and bicarbonate under acidic conditions SCN-(aq) + 2O2(aq) + 3H2O(l) = SO4-2(aq) + NH4+(aq) + HCO3-(aq) + H+(aq)   (eq. 11) Beyond reactions with sulphur and oxygen, cyanide can also form complexes with metal ions including iron (Ciminelli V. S., 1987; Farley, 1998; Jones, 1998; Marsden & House, 2006; Yannopoulos, 1991). This is another potential method of cyanide loss from leaching solutions, and so precipitation of excess iron from solution may be considered before the gold leaching stage. However, while ferrous iron-cyanide complexes may be found in certain cases, ferric iron complexation with cyanide is thought to be slow, and therefore not a major cause for cyanide loss in industry. Various iron-cyanide complexes are formed according to the reactions given below (Farley, 1998; Jones, 1998; Yannopoulos, 1991): Fe+2(aq) + 2CN-(aq) = Fe(CN)2(aq)        (eq. 12) Or: Fe+2(aq) + 3CN-(aq) = Fe(CN)3-(aq)        (eq. 13) Or: Fe+2(aq) + 6CN-(aq) = Fe(CN)6-4(aq)        (eq. 14) 12  Physical contact between the bulk solution containing cyanide and the gold grains inside ores is crucial to extracting the desired amount of gold from minerals. As will be discussed shortly, physical contact between leaching reagents and gold is not always initially possible, and so steps must be taken to pre-treat ores before gold leaching occurs (Klein, 2010; Komnitsas & Pooley, 1989; Kondos, Deschenes, & Morrison, 1995; La Brooy, Linge, & Walker, 1994). 2.2.2 Gold Leaching Using Alternative Leaching Agents  Although gold leaching in the mining industry typically occurs using the cyanidation technique, other leaching agents are sometimes used. More countries around the world are setting increasingly stringent environmental standards, in some instances entirely banning the use of cyanide for mining operations (International Cyanide Management Institute, 2012; Logsdon, Hagelstein, & Mudder, 1999; Marsden & House, 2006; Young & Jordan, 1995), and so the promise of alternative leaching agents is a tantalizing one. Cyanide is also a highly dangerous and expensive substance, with safety, environmental, and economic pressures increasing the drive to limit or abolish the use of this compound. It is for this reason that alternative leaching agents are constantly being researched and developed with the hopes of one day replacing cyanidation as the conventional method of extracting precious metals, including gold. Several leaching chemicals, or “lixiviants”, have been reported in literature. A key factor in determining the viability of a lixiviant for use in the gold extraction process is its stability. When gold forms a complex with a ligand, this aqueous compound must be stable enough to survive leaching and downstream conditions until the gold recovery stages, where the aqueous gold is concentrated and reconstituted by various means back into a solid form. The stability data of various gold complexes are given below in Table 2 (Aylmore, Alternative Lixiviants to Cyanide for Leaching Gold Ores, 2005; Aylmore & Muir, Thiosulfate Leaching of Gold - A Review, 2001; Deschenes, 2005; Senanayake, Gold Leaching in Non-Cyanide Lixiviant Systems: Critical Issues on Fundamentals and Applications, 2004; Swaminathan, Pyke, & Johnston, 1993; Yannopoulos, 1991).      13  Table 2: Averaged Stability Data for Select Gold Complexes Leaching Au+ Leaching Au+3 Lixiviant Complex Formula Stability Constant (β2) Lixiviant Complex Formula Stability Constant (β4) Cyanide Au(CN)2- 2.00∙1038 Cyanide Au(CN)4- 3.33∙1084 Thiosulphate Au(S2O3)2-3 3.37∙1028    Thiourea Au[CS(NH2)2]+ 1.67∙1023    Thiocyanate Au(SCN)2- 1.28∙1017 Thiocyanate Au(SCN)4- 3.18∙1043 Ammonia Au(NH3)2+ 5.00∙1025 Ammonia Au(NH3)4+ 1.00∙1048 Chloride AuCl2- 1.88∙109 Chloride AuCl4- 5.59∙1025 Bromide AuBr2- 1.00∙1012 Bromide AuBr4- 3.65∙1032 Iodide AuI2- 2.80∙1019 Iodide AuI4- 5.01∙1047  It is easy to see that gold cyanide complexes are the most stable among all researched lixiviants. This stability, coupled with the wide range of stability shown in the Pourbaix diagram for the gold-cyanide-water system, is why cyanidation is so widely used to leach gold. Another important factor that separates cyanidation from other lixiviants is the spontaneous dissolution of native (solid) gold into solution without application of electrochemical potential (Deschenes, 2005; La Brooy, Linge, & Walker, 1994; Swaminathan, Pyke, & Johnston, 1993). However, as mentioned, economic and environmental protection pressures from the public and governments of gold-producing nations have driven many companies in the precious metals sector to look for an alternative to cyanide. Among the most commonly researched and implemented lixiviants are thiosulphate (S2O3-2), thiourea (CS(NH2)2), ammonia (NH3(aq)), and various halides, including chloride (Cl-) (Aylmore, “Alternative Lixiviants…”, 2005; Farley, 1998; Klein, 2010; La Brooy, Linge, & Walker, 1994; Marsden & House, 2006; Prasad, Mensah-Biney, & Pizarro, 1991; Senanayake, “Gold Leaching…”, 2004; Swaminathan, Pyke, & Johnston, 1993; Yannopoulos, 1991). These compounds will be discussed below. 2.2.2.1 Gold Leaching Using Thiosulphate  The most viable lixiviant currently being researched as an alternative to cyanide is thiosulphate (S2O3-2). Thiosulphate is an intermediate sulphur species and is “metastable” meaning that it is thermodynamically unstable compared to elemental sulphur, sulphide (S-2), and sulphate (SO4-2) in solution. The idea of using thiosulphate to leach gold has been explored extensively, with increasing interest from the gold mining 14  industry as the use of cyanide is restricted in many gold-producing nations due to concerns over environmental toxicity and hazards to wildlife and human health (Aylmore & Muir, Thiosulfate Leaching of Gold - A Review, 2001). Thiosulphate will eventually be hydrolyzed to various polythionates including trithionate (S3O3-2), tetrathionate (S4O6-2), to sulphite (SO3-2), and finally the stable sulphate ion (SO4-2) (Jones, 1998). However, this oxidation can be quite slow due to the limited solubility of dissolved oxygen in solution, especially with leaching solutions with high ionic strengths or high temperatures (Aylmore, Alternative Lixiviants to Cyanide for Leaching Gold Ores, 2005). The chemistry of thiosulphate is quite complex, especially in copper-ammonia systems, and in the presence of various additives that either hasten the kinetics of gold complexation, prevent passivation of gold due to precipitation of surface layers, or increase the yield of precious metals from refractory ores. However, the basic idea remains relatively simple: gold is oxidized by a suitable oxidant in the presence of thiosulphate, which then forms a complex with the ligand, while the other half of the “redox” reaction is completed by the reduction of water to produce hydroxide. This is shown by the equations below (Zhang & Nicol, An Electrochemical Study of the Dissolution of Gold in Thiosulfate Solutions Part 1: Alkaline Solutions, 2003; Zhang & Nicol, An Electrochemical Study of the Dissolution of Gold in Thiosulfate Solutions Part II: Effect of Copper, 2005): Reduction: O2(aq) + 2H2O(l) + 4e- = 4OH-(aq)      (eq. 15a) Oxidation: Au(s) + 2S2O3-2(aq) = Au(S2O3)2-3 + e-      (eq. 15b) (1)(a) + (4)(b) Overall Reaction: 4Au(s) + 8S2O3-2(aq) + O2(aq) + 2H2O(l) = 4Au(S2O3)2-3(aq) + 4OH-(aq) (eq. 15) The above reaction has found to be slow, due to the degradation of thiosulphate and subsequent passivation of gold surfaces caused by thiosulphate and polythionates on the surface of gold particles (Zhang & Nicol, An Electrochemical Study of the Dissolution of Gold in Thiosulfate Solutions Part II: Effect of Copper, 2005). The various decomposition reactions for thiosulphate to other thionic species are given below (Aylmore & Muir, Thiosulfate Leaching of Gold - A Review, 2001): Oxidation of thiosulphate to tetrathionate occurring at pH 4 – 6: S2O3-2(aq) + O2(aq) + 2H2O(l) = 2SO4-2(aq) + 4H+(aq)      (eq. 16) Or: 4 S2O3-2(aq) + O2(aq) + 2H2O(l) = 2S4O6-2(aq) + 8OH-(aq)     (eq. 17) Disproportionation of thiosulphate directly into sulphate and native sulphur: 2S2O3-2(aq) + H2O(l) = 2SO4-2(aq) + 4S0(s) + OH-(aq)      (eq. 18) 15  Disproportionation of thiosulphate to sulphite: 3S2O3-2(aq) + 6OH-(aq) = 4SO3-2(aq) + 2S-2(aq) + 3H2O(l)     (eq. 19) Or: S2O3-2(aq) = SO3-2(aq) + S0(s)        (eq. 20) The disproportionation of thiosulphate to sulphite is not very likely to occur; thiosulphate may form hydrogen sulphite (HSO3-(aq)) and sulphur (S(s)) in water. The degradation of thiosulphate is an ongoing issue for gold leaching operations. To date, only one major gold mining operation has been implemented the use of thiosulphate to leach gold from mined ores: the Goldstrike project, currently owned by Barrick Gold Corporation (Barrick Gold Corporation, 2017). The success of this operation will no doubt spur on other companies to follow suit and utilize thiosulphate on a much larger commercial scale. 2.2.2.2 Gold Leaching Using Thiourea  Thiourea (CS(NH2)2) is another promising lixiviant that has enjoyed extensive study in the past. However, in recent years it has fallen out of favour with operators with the development of thiosulphate as a lixiviant, which is now seen as the main alternative to cyanide (Aylmore, “Alternative Lixiviants…”, 2005; Farley, 1998). The half-reaction for gold leaching using thiourea is given below (La Brooy, Linge, & Walker, 1994; Prasad, Mensah-Biney, & Pizarro, 1991; Swaminathan, Pyke, & Johnston, 1993; Yannopoulos, 1991): Au(s) + 2CS(NH2)2(aq) = Au[CS(NH2)2]2+(aq) + e-   E° = -0.38 V  (eq. 21) In the presence of ferric or other oxidizing agents, it is thought that thiourea forms an intermediate compound called “formamidine disulphide” (NH2(NH)CSSC(NH)NH2(aq)), which itself is a strong oxidant that can extract gold (Farley, 1998; Yannopoulos, 1991). The unbalanced reactions for the formation of formamidine disulphide and subsequent dissolution of gold are shown below (Yannopoulos, 1991): Step 1: 2CS(NH2)2(aq) = NH2(NH)CSSC(NH)NH2(aq) + 2H+(aq) + 2e-             E° > 0.30 V  (eq. 22a) Step 2: 2Au(s) + 2CS(NH2)2(aq) + NH2(NH)CSSC(NH)NH2(aq) + 2H+(aq) = 2Au[CS(NH2)2]2+          E° = 0.04 V  (eq. 22b) The low potential given for Eq. 22b above suggests that thiourea and formamidine disulphide are not strong oxidants. However, it should be noted that to achieve this low gold leaching potential, the oxidation of thiourea to formamidine disulphide must first occur at anodic potentials higher than 0.30V SHE. Equation 22b is a combination of a complexation reaction and a redox half-reaction. Thiourea leaching of gold and silver typically occurs in acidic media and has shown both high extraction levels of precious metals, and much faster kinetics than those seen with conventional cyanidation (Swaminathan, Pyke, & Johnston, 1993; 16  Yannopoulos, 1991). Some other benefits to using thiourea are that this compound is very selective towards gold and silver, and does not readily form complexes with base metals, unlike cyanide; additionally, thiourea is very effective in treating oxidized refractory sulphide ores as well as refractory carbonaceous ores (Farley, 1998; Prasad, Mensah-Biney, & Pizarro, 1991; Yannopoulos, 1991) However, it is hampered by a prohibitively high cost, coupled with substantial consumption of reagent through its oxidation (La Brooy, Linge, & Walker, 1994; Prasad, Mensah-Biney, & Pizarro, 1991; Swaminathan, Pyke, & Johnston, 1993). Additionally, thiourea is not as widely used in industry due to some safety and environmental concerns that this lixiviant may be hazardous to health as a carcinogen (Farley, 1998). 2.2.2.3 Gold Leaching Using Ammonia  The use of ammonia to leach gold has been researched for several decades, however very few ammoniacal processes have been implemented in industry (Aylmore, “Alternative Lixiviants…”, 2005). The chemistry of ammonia leaching is given below (Dasgupta, Guan, & Han, 1997; Guan & Han, 1996): Au(s) = Au+(aq) + e-      E° = 1.69 V  (eq. 23a) Au+(aq) + 2NH3(aq) = Au(NH3)2+(aq)       (eq. 23b) Overall Reaction: Au(s) + 2NH3(aq) = Au(NH3)2+(aq) + e-  E° = 0.572 V  (eq. 23) Normally, leaching occurs in a copper-ammonia solution at higher pH (~8.5 – 9.5) in order to avoid dissociation of ammonia into ammonium (NH4+) and is performed at elevated temperatures, (above 75°C, usually 100 – 300°C) and higher pressures (600 – 1000 kPa) to increase reaction kinetics of leaching (Dasgupta, Guan, & Han, 1997; Guan & Han, 1996; La Brooy, Linge, & Walker, 1994). The use of ammonia is often used to improve kinetics of gold dissolution and is associated with the more successful leaching of gold using thiosulphate (e.g. in copper-ammonia solutions). In a few instances, ammonia is used with halides such as iodine (I2) to oxidize gold ores and form gold-ammonia complexes (Aylmore, “Alternative Lixiviants…”, 2005). The use of halides to directly leach gold is discussed below. 2.2.2.4 Gold Leaching Using Halides  The use of halides to leach gold has been practiced since the late 1800s and was the method of choice before the advent of cyanidation (Farley, 1998; La Brooy, Linge, & Walker, 1994; Yannopoulos, 1991). Halogen anions (or halides) such as chloride (Cl-), bromide (Br-), and iodide (I-) have all been used to varying extents to complex gold in acidic media (Farley, 1998; Prasad, Mensah-Biney, & Pizarro, 1991; Senanayake, Gold Leaching in Non-Cyanide Lixiviant Systems: Critical Issues on Fundamentals and Applications, 2004). In many instances, halide leaching is used as a complementary method of gold leaching, either as a separate 17  step prior to or after conventional cyanidation, or as a co-lixiviant in the leaching stage, as with ammonia, thiosulphate, or thiourea. The general reaction for the dissolution of gold with a halide ion is given below (La Brooy, Linge, & Walker, 1994; Senanayake, Gold Leaching in Non-Cyanide Lixiviant Systems: Critical Issues on Fundamentals and Applications, 2004): Au(s) + nX-(aq) = AuXn-(n-m)(aq) + me-       (eq. 24) Where:  n is the stoichiometric factor associated with the halide ion,  X is a halogen ion (e.g. chloride (Cl-), bromide (Br-), iodide (I-), etc.)  m is the number of electrons released during the oxidation of gold The simplified general reactions for halide complexation with gold are shown below (Aylmore, Alternative Lixiviants to Cyanide for Leaching Gold Ores, 2005): 2Au(s) +3X2(g) + 2X-(aq) = 2AuX4-(aq) where X is Cl or Br    (eq. 25a) 2Au(s) + Y3-(aq) + Y-(aq) = 2AuY2-(aq) where Y is I     (eq. 25b) Generally, the observed order of dissolution rates for gold using halides is RCl->RBr->RI-, however the order of stability constants measured for these complexes is reversed: β2Cl-<βBr-<βI- (Aylmore, Alternative Lixiviants to Cyanide for Leaching Gold Ores, 2005; Qi & Hiskey, 1991; Senanayake, Gold Leaching in Non-Cyanide Lixiviant Systems: Critical Issues on Fundamentals and Applications, 2004). Halides are seen as a promising field for supplementary gold leaching, however tight control over leaching conditions must be maintained, and leaching of gold does not occur at any appreciable rate unless a substantial concentration of halide ions is maintained in solution – a difficult feat given that species such as native sulphur, sulphides, iron, and other base metals consume halogens by acting as reductants to destroy gold-halide complexes by reducing gold back into its solid form (Aylmore, Alternative Lixiviants to Cyanide for Leaching Gold Ores, 2005; La Brooy, Linge, & Walker, 1994; Yannopoulos, 1991). An example of this is shown below: AuCl4(aq) + S-2(aq) + 4H2O(l) = Au(s) + SO4-2(aq) + 4HCl(aq)     (eq. 26)  2.3 Processing of Refractory Gold Ores  “Refractory” gold ores are those that are difficult to process by conventional means, and therefore require extra steps or different processes to extract the precious metal from them (Klein, 2010; Komnitsas & Pooley, 18  1989; Marsden & House, 2006). Refractory ores are classified into one of three main categories: carbonaceous ores, silicate ores, and sulphide ores Refractory ores are those that have one ore more mineralogical types that fall under these categories, with some ores that can be classified as “double-“ or even “triple-refractory”, because of their mixed compositions (Klein, 2010; Komnitsas & Pooley, 1989; Kondos, Deschenes, & Morrison, 1995; Marsden & House, 2006). Each of the categories of refractory ores act in different ways to reduce the overall recovery of gold from ores. These effects are summarized below (Komnitsas & Pooley, 1989; La Brooy, Linge, & Walker, 1994; Marsden & House, 2006; Yannopoulos, 1991) : - Occlusion of finely disseminated (sometimes called “invisible”) gold. This is usually seen with sulphide or silicate ore types - Reaction between impurities in the ore with cyanide, resulting in cyanide consumption/degradation in the bulk solution. Ores with this problem are known as “cyanicides” - Reaction between impurities in the ore with oxygen, leading to oxygen consumption from the bulk solution. Ores that consume excess oxygen are known as “reactive” ores - Adsorption of gold complexed ligands from the “pregnant leach solution” (or PLS) onto impurities found on or within mineral particles. These ores are known as “preg-robbing” ores Various methods exist to deal with each of these problems, and these processes are currently being employed with generally high levels of success in the industry. Solutions for dealing with refractory ores include Ultra-fine grinding, Roasting, Bio-oxidation, Pressure Oxidation, and Chlorination, among many other processes (Afenya, 1991; Klein, 2010; Komnitsas & Pooley, 1989). For the sake of brevity, only techniques used for the pre-treatment of refractory iron sulphide minerals will be discussed here. These include Roasting, Bio-oxidation, and Aqueous Pressure Oxidation. 2.3.1 Roasting of Iron Sulphide Minerals  Roasting is the process where intense heat is applied to sulphide mineral concentrates to oxidize sulphides (S-2) and native sulphur (S8) in mineral ores (Fraser, Walton, & Wells, 1991; Klein, 2010; Komnitsas & Pooley, 1989). Roasting can be classified as a purely pyrometallurgical process, using oxygenated roasters and calcines to pyrolyze (or burn) sulphur from the ore, creating sulphur dioxide (SO2) or sulphur trioxide (SO3) gas. In more recent times, roasting has been performed in Fluidized Bed Reactors (FBRs) (Fraser, Walton, & Wells, 1991; Komnitsas & Pooley, 1989; Yannopoulos, 1991). Fluidized bed-reactors are 19  commonly used to introduce heat and oxygen into high-sulphur ore concentrates and these processes take place at anywhere from 600 – 800°C (Fraser, Walton, & Wells, 1991; Komnitsas & Pooley, 1989). Both iron and sulphur are oxidized, but to different extents, with sulphur generally being the more reacted species. Common iron and sulphur oxidation reactions are given below (Komnitsas & Pooley, 1989; La Brooy, Linge, & Walker, 1994; Marsden & House, 2006): Oxidation of pyrite/marcasite by oxygen, producing hematite 4FeS2(s) + 11O2(aq) = 2Fe2O3(s) + 8SO2(g)       (eq. 27a) Or: 3FeS2(s) + 8O2(aq) = Fe3O4(s) + 6SO2(g)       (eq. 28a) 4Fe3O4(s) + O2(aq) = 6Fe2O3(s)        (eq. 28b) (4)(a) + (1)(b) 12FeS2(s) + 33O2(aq) = 6Fe2O3(s) + 24SO2(g)      (eq. 28) Oxidation of arsenopyrite by oxygen, producing hematite and arsenic trioxide Step 1: 12FeAsS(s) + 29O2(aq) = 4Fe3O4(s) + 6As2O3(s) + 12SO2(g)    (eq. 29a) Step 2: 4Fe3O4(s) + O2(aq) = 6Fe2O3(s)       (eq. 29b) Overall Reaction: 2FeAsS(s) + 5O2(aq) = Fe2O3(s) + As2O3(s) + 2SO2(g)   (eq. 29) The Roasting process is well studied and is still implemented widely throughout the world (Fraser, Walton, & Wells, 1991; Komnitsas & Pooley, 1989; La Brooy, Linge, & Walker, 1994; Marsden & House, 2006). However, in recent years, environmental pressures have forced many mines and processing plants to consider other options, due to the excessive air pollution that is emitted (Fraser, Walton, & Wells, 1991; Klein, 2010; Komnitsas & Pooley, 1989; La Brooy, Linge, & Walker, 1994). This is especially true as more ore concentrates today are of poorer quality, with arsenic, mercury, and lead impurities present among many other toxic and harmful pollutants that might be volatilized into the air during roasting. The capital cost associated with Roasting results from the need to purify vented off-gas emitted from the roasters, and high operating costs stem from the energy demand required to run a high temperature process to remove sulphur from mineral concentrates (Komnitsas & Pooley, 1989). 2.3.2 Bio-oxidation of Iron Sulphide Minerals  The process of using bacteria to oxidize sulphide minerals has been in industrial use since the early 1990s (Komnitsas & Pooley, 1989; La Brooy, Linge, & Walker, 1994; Marsden & House, 2006). Bio-oxidation plants are increasing in popularity around the world, as the various operational aspects of the process are more thoroughly understood by process operators. In this process, various types of bacteria can be used to 20  oxidize either iron, or native sulphur and sulphides found in particles of mineral concentrates. The most commonly used bacteria are Thiobacillus Ferrooxidans, Thiobacillus Thiooxidans, Sulpholobus, and Leptospirillum Ferrooxidans (Jones, 1998; Komnitsas & Pooley, 1989; Marsden & House, 2006). Numerous commonly occurring reactions for bio-oxidation of iron sulphide minerals are given below (Jones, 1998; Komnitsas & Pooley, 1989; Marsden & House, 2006; Yannopoulos, 1991): Bacterial assisted oxidation of pyrite 2FeS2(s) + 7O2(aq) + 2H2O(l) bacteria = 2FeSO4(aq) + 2H2SO4(aq)    (eq. 30) Or: 2FeS2(s) + O2(aq) + 2H2SO4(aq) bacteria = 2FeSO4(aq) + 4S0(s) + 2H2O(l)    (eq. 31) Bacterial assisted oxidation of arsenopyrite 2FeAsS(s) + 7O2(aq) + 2H2O(l) + H2SO4(aq) bacteria = 2H3AsO4(aq) + Fe2(SO4)3(aq)  (eq. 32) Or: 2FeAsS(s) + 7O2(aq) + 5H2O(l) bacteria = Fe2O3(s) + 2 H2SO4(aq) + 2 H3AsO4(aq)  (eq. 33) Or: 4FeAsS(s) + 13O2(aq) + 6H2O(l) bacteria = 4H3AsO4(aq) + 4FeSO4(aq)    (eq. 34) Or: 4FeAsS(s) + 11O2(aq) + 6H2O(l) bacteria = 4H3AsO3(aq) + 4FeSO4(aq)    (eq. 35) Oxidation of ferrous- to ferric sulphate 4FeSO4(aq) + O2(aq) + 2H2SO4(aq) bacteria = 2Fe2(SO4)3(aq) + 2H2O(l)    (eq. 36) Oxidation of arsenic (III) acid to arsenic (V) acid 2H3AsO3(aq) + O2(aq) = 2H3AsO4(aq)       (eq. 37) Oxidation of arsenopyrite by ferric sulphate 4FeAsS(s) + 4Fe2(SO4)3(aq) + 3O2(aq) + 6H2O(l) = 4H3AsO3(aq) + 12FeSO4(aq) + 4S0(s)  (eq. 38) Or: 2FeAsS(s) + 2Fe2(SO4)3(aq) + 5O2(aq) + 3H2O(l) = 2H3AsO4(aq) + 6FeSO4(aq) + 2S0(s)  (eq. 39) Bacterial assisted oxidation of sulphur 2S0(s) + 3O2(aq) + 2H2O(l) bacteria = 2H2SO4(aq)      (eq. 40) Many of the reactions shown above occur under slightly different conditions from each other, and these processes have been studied extensively over at least the past three decades (Jones, 1998). Most bio-oxidation occurs in low pH conditions (0.5 – 3.0) and at ambient or slightly elevated temperatures (25 – 21  80°C), depending on which bacteria are used in the oxidation process (Fraser, Walton, & Wells, 1991; Jones, 1998; Marsden & House, 2006; Yannopoulos, 1991). Both Capital and Operating costs associated with bio-oxidation are much lower than other types of oxidation, and it is for this reason that bio-oxidation is increasingly more attractive as a pre-treatment option for refractory ores (Fraser, Walton, & Wells, 1991; Jones, 1998; Komnitsas & Pooley, 1989). However, drawbacks of this process include a slow oxidation time in the order of several days for iron or sulphur to completely react inside a bio-oxidation tank, stringent requirements and control of operating conditions (including limits on the concentrations of aqueous base metals that may also be released during the oxidation process), as well as problems for equipment and process stages downstream (e.g. corrosion of equipment and fouling of activated carbon in the gold loading and elution stage) (Fraser, Walton, & Wells, 1991; Jones, 1998; La Brooy, Linge, & Walker, 1994). Another important consideration for bacterial oxidation is that this process generates reactive sulphur species such as thiosulphate (S2O3-2), sulphite (SO3-2). The reason for this is generally thought to be incomplete oxidation of sulphides (S-2) in mineral ore concentrates, or production of biological by-products (i.e. enzymes) which might catalyze reaction between thiosulphate and cyanide ions (Jones, 1998; Miller & Brown, 2005). Sulphides are oxidized via oxygen or ferric (Fe+3) to produce native sulphur, or various thionate compounds (e.g. trithionate S3O6-2, or tetrathionate S4O6-2). These compounds are then further oxidized to sulphate (SO4-2). Examples of some of these reaction are shown below (Jones, 1998; Sand, Gehrke, Jozsa, & Schippers, 2001; Schippers, Rohwerder, & Sand, 1999; Schippers & Sand, Bacterial Leaching of Metal Sulfides Proceeeds by Two Indirect Mechanisms via Thiosulfate or via Polysulfides and Sulfur, 1999; Schippers, Jozsa, & Sand, Sulfur Chemistry in Bacterial Leaching of Pyrite, 1996): Sulphide (S-2) to native sulphur (S0) to sulphate (SO4-2) Step 1: FeS2(s) + 2Fe+3(aq) = 3Fe+2(aq) + 2S0(s)      (eq. 41a) Step 2: 2S0(s) + 3O2(aq) + 2H2O(l) = 2SO4-2(aq) + 4H+(aq)     (eq. 41b) Overall Reaction: FeS2(s) + 2Fe+3(aq) + 3O2(aq) + 2H2O(l) = 3Fe+2(aq) + 2SO4-2(aq) + 4H+(aq) (eq. 41) Sulphide (S-2) to thiosulphate (S2O3-2) to sulphate (SO4-2) Step 1: 2S-2(s) + 2O2(aq) + H2O(l) = S2O3-2(aq) + 2OH-(aq)     (eq. 42a) Step 2: S2O3-2(aq) + O2(aq) + OH-(aq) = 2SO4-2(aq) + H2O(l)     (eq. 42b) Overall Reaction: S-2(s) + 3O2(aq) = OH-(aq) + 2SO4-2(aq)     (eq. 42) Or: Step 1: FeS2(s) + 6Fe+3(aq) + 3H2O(l) = S2O3-2(aq) + 7Fe+2(aq) + 6H+(aq)   (eq. 43a) 22  Step 2: S2O3-2(aq) + 8Fe+3(aq) + 5H2O(l) = 2SO4-2(aq) + 8Fe+2(aq) + 10H+(aq)   (eq. 43b) Overall Reaction: FeS2(s) + 14Fe+3(aq) + 8H2O(l) = 2SO4-2(aq) + 15Fe+2(aq) + 16H+(aq)  (eq. 43) Sulphite (SO3-2) to sulphate (SO4-2) 2SO3-2(aq) + O2(aq) = 2SO4-2(aq)        (eq. 44) The reactions above are just a few of the possible reactions by which sulphur in its various forms can be oxidized to various thionic compounds. The overall order of sulphur reduction however remains consistent, with a step-wise oxidation of sulphur from an oxidation state of zero to an oxidation state of +6. The simplified order of oxidation is as follows (Jones, 1998): S-2 → S0/S8 → S2O3-2 → SO3-2 → SO4-2 However, many of the above intermediate reactions may be ignored or disregarded in this body of research, due to several key factors. First, the experiments undertaken for this work were done under mainly alkaline conditions, whereby any acidic species (e.g. H+) would react immediately with alkaline species such as hydroxide (OH-), and some reactions, such as those involving H+ ions, may even be prevented from forming due to the high pH of the system. The stability of sulphur compounds, including intermediate or “metastable” thionic complexes such as thiosulphate or sulphite, can be illustrated using Pourbaix diagrams for the stable and metastable S-H2O systems at ambient conditions, given below in Figure 7 and Figure 8, respectively: 23   Figure 7: Pourbaix diagram for the thermodynamically stable S-H2O system at atmospheric conditions and 25°C with 0.1 M sulphur concentration by HSC 6.0  Figure 8: Pourbaix diagram for the thermodynamically metastable S-H2O system at atmospheric conditions and 25°C with 0.1 M sulphur concentration, and excluding some species such as sulphate, trithionate, and pentathionate, by HSC 6.0 24  Secondly, the concentrations of intermediate thionic compounds such as trithionate or tetrathionate are difficult to determine in situ at any point in time of an experiment. This is due to the instability of these species driving the formation of more stable compounds such as sulphate. Lastly, in a system with an abundance of oxygen, it may be assumed that all intermediate aqueous sulphur species may be oxidized further to sulphate. It was initially hypothesized by this author that, because of the abundant dosage of air introduced into the leaching reactor over the course of all tests performed, any sulphur oxidation that occurred in the system was driven to completion over the course of eight hours. However, from the results of thiocyanate testing, this hypothesis was disproven by the fact that partially oxidized aqueous thionic species in solution reacted with sodium cyanide to produce significant levels of thiocyanates. At the start of this research, it was assumed that all sulphur found as a solid was in the form of either native sulphur or as part of sulphide minerals, and any sulphur found in solution was in the form of sulphate at the end of the tests. The presence of trace amounts of other aqueous sulphur species was initially not anticipated to have a significant effect on results. However, for all oxidation calculations, aqueous sulphur concentrations were used to compute total oxidation and model oxidation levels over time.  2.3.3 Aqueous Pressure Oxidation of Iron Sulphide Minerals  The use of aqueous Pressure Oxidation of refractory ores is widespread in the precious metals industry, and it is generally seen to have the greatest degree of success in ore dissolution and subsequent gold leaching, especially for those with more than one type of refractory mineral associated with gold (Fraser, Walton, & Wells, 1991; Papangelakis, “Aqueous Pressure Oxidation…”, 1986; Prasad, Mensah-Biney, & Pizarro, 1991). Pressure Oxidation works in much the same way that Atmospheric Oxidation works, except that intense heat and pressures can be used to force chemical reactions to happen that would normally never occur in ambient conditions. Typically, the conditions inside a leaching pressure reactor, or autoclave, reach 90 – 250°C, and upwards of 1300 – 3500 kPa (200 – 500 psi) (Koslides & Ciminelli, 1992; Papangelakis, Aqueous Pressure Oxidation of Arsenopyrite, 1986; Papangelakis & Demopoulos, Acid Pressure Oxidation of Arsenopyrite: Part I, Reaction Chemistry, 1990; Papangelakis & Demopoulos, Acid Pressure Oxidation of Pyrite: Reaction Kinetics, 1991; Prasad, Mensah-Biney, & Pizarro, 1991). The reason for this is that higher temperatures and pressures contribute thermal energy that is used to overcome activation energies for kinetically slow reactions; these conditions also allow the electrochemical potential of the system to increase, so that oxidation reactions with higher potentials can occur. Given below are just some of the many reactions that occur in pressure oxidation of iron sulphide minerals (Amoah-Forson, 1986; Fleming, Basic Iron Sulfate - A Potential Killer in the Processing of Refractory Gold Concentrates by Pressure Oxidation, 2010; Koslides & Ciminelli, 1992; Papangelakis, Aqueous Pressure 25  Oxidation of Arsenopyrite, 1986; Papangelakis & Demopoulos, Acid Pressure Oxidation of Arsenopyrite: Part I, Reaction Chemistry, 1990; Papangelakis & Demopoulos, Acid Pressure Oxidation of Arsenopyrite: Part II, Reaction Kinetics, 1990; Papangelakis & Demopoulos, Acid Pressure Oxidation of Pyrite: Reaction Kinetics, 1991; Thomas, 2005) (Zhang S. , Oxidation of Refractory Gold Concentrates and Simultaneous Dissolution of Gold in Aerated Alkaline Solutions, 2004): In acidic conditions: Oxidation of pyrite (under medium temperatures/pressures) 2FeS2(s) + 7O2(aq) + 2H2O(l) = 2FeSO4(aq) + 2H2SO4(aq)     (eq. 45) Or: 4FeS2(s) + 15O2(aq) + 2H2O(l) = 2Fe2(SO4)3(aq) + 2H2SO4(aq)     (eq. 46) “Incomplete” oxidation of pyrite, leaving native sulphur as a by-product FeS2(s) + 2O2(aq) = FeSO4(aq) + S0(s)       (eq. 47) Or: FeS2(s) + Fe2(SO4)3(aq) = 3FeSO4(aq) + 2S0(s)      (eq. 48) Oxidation of arsenopyrite (under medium temperatures/pressures) 4FeAsS(s) + 11O2(aq) + 2H2O(l) = 4FeSO4(aq) + 4HAsO2(aq)     (eq. 49) Or: 4FeAsS(s) + 5O2(aq) + 4H2SO4(aq) = 4FeSO4(aq) + 4HAsO2(aq) + 4S0(s) + 2H2O(l)  (eq. 50) Or: 2FeAsS(s) + 7Fe2(SO4)3(aq) + 8H2O(l) = 16FeSO4(aq) + 2H3AsO4(aq) + 5H2SO4(aq) + 2S0(s) (eq. 51) Or: 4FeAsS(s) + 13O2(aq) + 2H2SO4(aq) + 2H2O(l) = 2Fe2(SO4)3(aq) + 2H3AsO4(aq) + 2HAsO2(aq) (eq. 52) Oxidation of arsenopyrite to produce hematite, arsenic acid (H3AsO4(aq)), and sulphuric acid (at high temperatures and pressures) 2FeAsS(s) + 7O2(aq) + 5H2O(l) = Fe2O3(s) + 2H3AsO4(aq) + 2H2SO4(aq)   (eq. 53) Formation of jarosite from hematite (at high temperatures and pressures) 3Fe2O3(s) + 4H2SO4(aq) + 5H2O(l) = 2(H3O·Fe3(SO4)2·(OH)6)(s)    (eq. 54) Oxidation of native sulphur 2S0(s) + 3O2(aq) + 2H2O(l) = 2H2SO4(aq)       (eq. 55) Or: S0(s) + 3Fe2(SO4)3(aq) + 4H2O(l) = 6FeSO4(aq) + 4H2SO4(aq)     (eq. 56) 26  Oxidation of ferrous to ferric- sulphate 4FeSO4(aq) + O2(aq) + 2H2SO4(aq) = 2Fe2(SO4)3(aq) + 2H2O(l)    (eq. 57) Oxidation of arsenic (III) acid to arsenic (V) acid 2HAsO2(aq) + O2(aq) + 2H2O(l) = 2H3AsO4(aq)      (eq. 58) Or: HAsO2(aq) + O2(aq) + 2FeSO4(aq) + H2SO4(aq) = Fe2(SO4)3(aq) + H3AsO4(aq)   (eq. 59) Precipitation of iron-oxides, - hydroxides, and -oxy-hydroxides also occur: Precipitation of hematite Fe2(SO4)3(aq) + 3H2O(l) = Fe2O3(s) + 3H2SO4(aq)      (eq. 60) Or: Fe2(SO4)3(aq) + 2H2O(l) = Fe2O3(s) + H2SO4(aq)      (eq. 61) Precipitation of jarosites 3Fe2(SO4)3(aq) + 14H2O(l) = 2H3OFe3(SO4)2(OH)6(s) + 5H2SO4(aq)    (eq. 62) Or: 3Fe2(SO4)3(aq) + 12H2O(l) + M2SO4(aq) = 2MFe3(SO4)2(OH)6(s) + 6H2SO4(aq)  (eq. 63) (Where M = Ag+, NH4+, K+, ½Pb+2, etc.) Precipitation of “scorodite” (or ferric arsenate) from aqueous arsenic acid and ferric sulphate 2H3AsO4(aq) + Fe2(SO4)3(aq) = 2FeAsO4(s) + 3H2SO4(aq)     (eq. 64) In alkaline conditions: 4FeS2(s) + 15O2(aq) + 16NaOH(aq) = 2Fe2O3(s) + 8Na2SO4(aq) + 8H2O(l)   (eq. 65) 2FeAsS(s) + 7O2(aq) + 10NaOH(aq) = Fe2O3(s) + 2Na3AsO4(aq) + 2Na2SO4(aq) + 5H2O(l) (eq. 66) Pressure oxidation has two main advantages over atmospheric oxidation: increased reaction kinetics and a greater extent of oxidation of refractory ores. The difference in overall leach time between an atmospheric leach and a pressurized process is significant. Atmospheric oxidation processes can take at least several hours, days, or, in the case of heap leaching, even months to fully oxidize mineral ores; pressure oxidation is typically completed within one to eight hours (Komnitsas & Pooley, 1989; Papangelakis, Aqueous Pressure Oxidation of Arsenopyrite, 1986). Pressure leach processes generally oxidize mineral ores to a greater extent than atmospheric processes (Papangelakis, Aqueous Pressure Oxidation of Arsenopyrite, 1986; Papangelakis & Demopoulos, Acid Pressure Oxidation of Pyrite: Reaction Kinetics, 1991; Thomas, 2005). This is because higher pressures provide an increase in the presence of gases in solution due to 27  Henry’s law (Dreisinger, 2014). Additionally, higher temperatures provide thermal energy which acts to increase the kinetics of leach reactions.  2.4 Physical and Chemical Characteristics of Iron Sulphide Minerals  There are myriad sulphide minerals that are typically associated with precious metals such as gold and silver (Kondos, Deschenes, & Morrison, 1995; Marsden & House, 2006; Yannopoulos, 1991). The main types of iron-sulphide minerals associated with gold are pyrite/marcasite (various forms of FeS2) and arsenopyrite (FeAsS). Pyrrhotite (FeS) is also a gold-bearing iron sulphide but is not as commonly associated with gold as pyrite and arsenopyrite. Other types of sulphides, such as copper- and zinc- sulphides are also found to bear gold in some cases, these include host minerals such as covellite (CuS). chalcocite (Cu2S), chalcopyrite (CuFeS), and galena (PbS), among many others (Marsden & House, 2006; Yannopoulos, 1991). For this thesis, the three most common gold bearing iron sulphides will be examined: pyrite, marcasite, and arsenopyrite. 2.4.1 Pyrite and Marcasite  Pyrite is the world’s most common sulphide mineral, and the most common mineral associated with gold (Ciminelli V. S., 1987; Zhang & Nicol, An Electrochemical Study of the Dissolution of Gold in Thiosulfate Solutions Part II: Effect of Copper, 2005; Zhu, Li, & Wadsworth, Characterization of Surface Layers Formed During Pyrite Oxidation, 1994). The molecular formula for pyrite is FeS2. Pyrite is sometimes referred to as iron disulphide indicating its constituent species: ferrous iron (Fe+2), and a di-sulphide molecule (S2-2) (Lowson, 1982; Uhlig, Szargan, Nesbitt, & Laajelehto, 2001). There are two main crystal structures associated with the chemical formula FeS2: face-centred cubic, and orthorhombic – dipyramidal (Lowson, 1982; Mishra & Osseo-Asare, 1988; Uhlig, Szargan, Nesbitt, & Laajelehto, 2001). The face-centered cubic crystal structure is isometric, uniform and is classified as “pyrite” (Lowson, 1982). Given below is a schematic of the crystal structure of pyrite (Uhlig, Szargan, Nesbitt, & Laajelehto, 2001): 28   Figure 9: Face centered cubic crystal structure of pyrite; small black spheres indicate iron, larger grey spheres indicate sulphur (Uhlig, Szargan, Nesbitt, & Laajelehto, 2001) Pyrite is a pale-yellow mineral with a metallic lustre, with an extremely similar appearance to gold. It is this visual similarity to gold that gives pyrite the moniker “Fool’s Gold”. The relevant physical characteristics of pyrite are given below (Lowson, 1982; Mineralogical Society of America, 2003): Table 3: Physical and Chemical Properties of Pyrite Property Value Unit Molecular Formula FeS2  Formula Weight 119.98 g/mol Crystal Structure Face-Centered Cubic (FCC)  Density 5014 kg/m3 Melting Point 1171 (1444) °C (K) Electrical Resistivity (20°C) 1∙10-3 (n-type), 2∙10-2 (p-type) Ω∙m Mohs Hardness 6 – 6.5  ΔGf°298K -166.94 kJ/mol  In contrast to pyrite, marcasite has an orthorhombic crystal structure, implying that the normally uniform cubic structure has been stretched in one or more directions (Lowson, 1982; Uhlig, Szargan, Nesbitt, & Laajelehto, 2001). The structure of marcasite is “di-pyramidal”; meaning that one plane of atoms in a crystal of marcasite is shared with its neighbour (creating two pyramids, stacked on each other). A schematic of the crystal structure of marcasite is given below in Figure 10 (Uhlig, Szargan, Nesbitt, & Laajelehto, 2001): 29   Figure 10: Face centered cubic crystal structure of marcasite; small black spheres indicate iron, larger grey spheres indicate sulphur (Uhlig, Szargan, Nesbitt, & Laajelehto, 2001) Marcasite is very similar to pyrite in appearance, with a pale brass-yellow or white colour, and a metallic lustre. Marcasite is considered to be more reactive than pyrite (Asta, Cama, & Acero, Dissolution Kinetics of Marcasite at Acidic pH, 2010; Papangelakis, Aqueous Pressure Oxidation of Arsenopyrite, 1986; Papangelakis & Demopoulos, Acid Pressure Oxidation of Arsenopyrite: Part I, Reaction Chemistry, 1990; Uhlig, Szargan, Nesbitt, & Laajelehto, 2001) The relevant physical characteristics of marcasite are given below (Lowson, 1982; Mineralogical Society of America, 2003): Table 4: Physical and Chemical Properties of Marcasite Property Value Unit Molecular Formula FeS2  Formula Weight 119.98 g/mol Crystal Structure Orthorhombic – Dipyramidal   Density 4887 kg/m3 Melting Point 450 (723) °C (K) Mohs Hardness 6 – 6.5  ΔGf°298K -158.57 kJ/mol  To understand the chemical breakdown of pyrite and marcasite, it is necessary to first examine a Pourbaix diagram of the stable Fe-H2O system and Fe-S-H2O system at ambient conditions. These are given below in Figure 11, and Figure 12, respectively: 30   Figure 11: Pourbaix diagram for the thermodynamically stable Fe-H2O system at 25°C with 0.1 mM iron concentration by HSC 6.0  Figure 12: Pourbaix diagram for the thermodynamically stable Fe-S-H2O system at 25°C with 0.1 mM for both iron and sulphur concentrations, by HSC 6.0 31  The only noticeable difference between the stable systems shown above and their metastable counterparts is the presence of goethite (FeOOH) in the upper ranges of electrochemical potential where hematite would be stable. Goethite is a metastable iron oxide hydroxide phase that generally transitions over a long period to a more stable hematite (Fe2O3) or magnetite (Fe3O4) phase once iron in solution precipitates. At a potential of 0V (SHE), Pyrite is generally seen to have a wide range of stability, from very low pH (~0.5) to around pH 10 or 11. However, in the presence of oxidants that increase the electrochemical potential (Eh) of the system (e.g. oxygen, hydrogen peroxide (H2O2), and ferric iron (Fe+3), pyrite and especially marcasite has been shown by numerous studies to be readily oxidized, releasing aqueous forms of iron and sulphur (Ahlberg, Forssberg, & Wang, 1990; Caldeira, Ciminelli, Dias, & Osseo-Asare, 2003; Ciminelli V. S., 1987; Zhang S. , Oxidation of Refractory Gold Concentrates and Simultaneous Dissolution of Gold in Aerated Alkaline Solutions, 2004). At lower electrochemical potentials, pyrite and marcasite oxidizes to create elemental sulphur (Ahlberg, Forssberg, & Wang, 1990; Mishra & Osseo-Asare, 1988). However, at higher anodic potentials sulphur is released into solution as thiosulphate, sulphite, and various other thionic compounds which are all eventually oxidized into the most stable aqueous sulphur product, sulphate (SO4-2) (Ahlberg, Forssberg, & Wang, 1990; Farley, 1998; Mishra & Osseo-Asare, 1988; Schippers, Rohwerder, & Sand, 1999; Schippers, Jozsa, & Sand, Sulfur Chemistry in Bacterial Leaching of Pyrite, 1996). A clear majority of hydrometallurgical studies focused on pyrite have attempted to either determine the reaction chemistry and mechanisms by which the oxidation of this gold bearing mineral occurs, approximate reaction kinetics for such reactions, or characterize the surface layers of iron and sulphur precipitated as oxidation by-products.  2.4.2 Arsenopyrite  Arsenopyrite is one of the most common iron sulphide minerals associated with gold, and has been researched by numerous authors over the past five decades (Bhakta, Langhans, & Lei, 1989; Ciminelli V. S., 1987; Espiell, Roca, Cruells, & Nunez, 1986; Papangelakis, Aqueous Pressure Oxidation of Arsenopyrite, 1986; Papangelakis & Demopoulos, Acid Pressure Oxidation of Arsenopyrite: Part I, Reaction Chemistry, 1990). Similar in composition to pyrite and marcasite, it is composed of iron, arsenic, and sulphur, with the arsenic substituting one sulphur anion in the crystal structure of the mineral. The molecular formula of arsenopyrite is FeAsS, but the exact ratio of arsenic to sulphur can vary between 0.9 to 1.1 for each element (Zhang S. , Oxidation of Refractory Gold Concentrates and Simultaneous Dissolution of Gold in Aerated Alkaline Solutions, 2004). Arsenopyrite is composed of a ferrous iron (Fe+2) ion bonded to a “realgar” – like dianion of AsS-2 (Papangelakis, Aqueous Pressure Oxidation of 32  Arsenopyrite, 1986; Tossel, Vaughan, & Burdett, 1981). The monoclinic crystal structure of arsenopyrite is thought to be extremely like that of marcasite, except that the sulphur di-anion normally associated with marcasite is modified slightly by replacing one sulphur atom with an arsenic atom (Lowson, 1982; Tossel, Vaughan, & Burdett, 1981). Arsenopyrite breaks down in oxidation by first releasing each of these ions into solution, and further oxidation of the AsS-2 ion decomposes into elemental arsenic and sulphide (S-2) ions, or As- and S- ions (Nesbitt, Muir, & Pratt, 1995; Zhang S. , Oxidation of Refractory Gold Concentrates and Simultaneous Dissolution of Gold in Aerated Alkaline Solutions, 2004). Literature has found that the AsS-2 complex is typically comprised of As- and S- ions (Nesbitt, Muir, & Pratt, 1995). Because of the compositional similarities between arsenopyrite and pyrite, and the fact that both minerals break down into ferrous iron and a sulphur-containing dianion, arsenopyrite is sometimes referred to in industry as “arsenical pyrite”. This implies that the arsenic found in the crystal structure of arsenopyrite is substituted in place of some sulphur ions, creating a corrupted form of regular pyrite (Lowson, 1982; Papangelakis, Aqueous Pressure Oxidation of Arsenopyrite, 1986). Some relevant physical and chemical characteristics of arsenopyrite are given below in Table 5 (Mineralogical Society of America, 2003; Smyth & McCormick, 1994): Table 5: Physical and Chemical Properties of Arsenopyrite Property Value Unit Molecular Formula FeAsS  Formula Weight 162.83 g/mol Crystal Structure Monoclinic   Density 6130 kg/m3 Mohs Hardness 5.5 – 6  ΔGf°298K -119.66 kJ/mol  To understand the reaction chemistry of arsenopyrite oxidation, it is necessary to first examine the Pourbaix diagrams for the As-H2O system, the As-S-H2O system, and the stable and metastable Fe-As-S-H2O systems. These diagrams are given below in Figure 13, Figure 14, Figure 15, and Figure 16, respectively: 33   Figure 13: Pourbaix diagram for the thermodynamically stable As-H2O system at 25°C with 0.1 mM for arsenic concentration, by HSC 6.0  Figure 14: Pourbaix diagram for the thermodynamically stable As-S-H2O system at 25°C with 0.1 mM for both arsenic and sulphur concentrations, by HSC 6.0 34   Figure 15: Pourbaix diagram for the thermodynamically stable Fe-As-S-H2O system at 25°C with 0.1 mM for iron, arsenic, and sulphur concentrations, by HSC 6.0  Figure 16: Pourbaix diagram for the thermodynamically metastable Fe-As-S-H2O system at 25°C with 0.1 mM for iron, arsenic, and sulphur concentrations, omitting the scorodite phase, by HSC 6.0 35  For Figure 16 above, the metastable scorodite phase is omitted, to show the boundary for ferric iron, which occurs at an Eh of roughly 0.75 V SHE. At a concentration of 0.1 mM aqueous iron, the region of stability for ferric iron ends at a pH of 1.75; this region increases with lowered iron concentrations. Ferric iron and pentavalent arsenic precipitate scorodite (FeAsO4·2H2O) at high electrochemical potentials and low pH. Combining the metastable systems above, a more complex Eh-pH diagram, which includes Scorodite, can be drawn, as given below in Figure 17 (Bhakta, Langhans, & Lei, 1989):  Figure 17: Complex Pourbaix diagram for the thermodynamically stable Fe-As-S-H2O system at 25°C with 0.1 mM for iron, arsenic, and sulphur concentrations (Bhakta, Langhans, & Lei, 1989).  For the diagram above, solid lines refer to areas of stability for the iron-arsenic-sulphur-water diagram and overlaid dashed lines refer to areas of stability determined by the arsenic-sulphur-water diagram, indicating a difference in arsenical acids or sulphur by-products evolved under specific conditions. As with pyrite and marcasite, arsenopyrite has a large area of stability across a wide range of pH values. However, at high pH (basic or alkaline conditions) arsenopyrite begins to be amenable to oxidation in the 36  presence of water at increasingly lower electrochemical potential. At a pH of 10, arsenopyrite is no longer stable at -0.5 V SHE, and is oxidized into ferric iron (Fe+3) and hydrogen arsenate (HAsO4-2) which includes highly oxidized arsenic (As (V)) in solution. Ferric iron and arsenate species can precipitate as many different species, including iron hydroxides, -oxy-hydroxides, -jarosites, goethite (usually as a metastable precursor to other iron oxide species), and scorodite (FeAsO4·nH2O), nearly all of which can serve to capture iron and arsenic from solution (jarosites do not normally capture arsenic, for example). These precipitates have different topologies and compositions, so it is important for process operators to select chemistries that will serve their needs, especially with regards to layer thickness and porosity, as well as settling properties in solid/liquid separation stages.  2.5 Atmospheric Oxidation of Iron Sulphide Minerals  The oxidation of iron sulphide minerals in atmospheric, low temperature conditions has been comprehensively studied by numerous studies over the past six decades. Initially, the mining industry considered using this type of oxidation to remove sulphur from uranium ores, as native sulphur and sulphides are very undesirable impurities, and hinder further purification of uranium due to the viscous nature of sulphur above a temperature of 159°C (Ciminelli V. S., 1987; Dreisinger, 2014). The coal industry has also considered oxidizing coal-bearing sulphide minerals to remove sulphur impurities, and due to stricter environmental regulations over the years, the application of Roasting has given way to hydrometallurgical leaching techniques to oxidize sulphide minerals of all types (Fan, Markuszewski, & Wheelock, 1984). Since then, the majority of research has focused on degradation of iron sulphide ores for the purpose of enhancing precious metals recovery (Farley, 1998; Fraser, Walton, & Wells, 1991; Prasad, Mensah-Biney, & Pizarro, 1991; Zhang S. , Oxidation of Refractory Gold Concentrates and Simultaneous Dissolution of Gold in Aerated Alkaline Solutions, 2004). Oxidation processes to treat iron sulphide minerals such as pyrite, marcasite, and arsenopyrite are generally divided between the use of acidic, or alkaline media. Ciminelli noted in her summary of literature that most of the research conducted over the past fifty years has examined acidic media (Asta, Cama, & Acero, Dissolution Kinetics of Marcasite at Acidic pH, 2010; Lowson, 1982; Nesbitt, Muir, & Pratt, 1995; Tao, Li, Richardson, & Yoon, 1994; Zhu, Li, & Wadsworth, Characterization of Surface Layers Formed During Pyrite Oxidation, 1994). This may be explained by the higher extents of oxidation and rates of reaction generally observed with acid dissolution compared to processes using alkaline media (Ciminelli V. S., 1987; Farley, 1998). 37  However, despite the slower kinetics and complications arising from precipitation of iron- and sulphur- surface layers, various studies have indeed measured the efficacy of reactions of alkaline processes for the purposes of pre-treating mineral ores in this manner in a full-scale precious metals recovery operation (Ahlberg, Forssberg, & Wang, 1990; Bhakta, Langhans, & Lei, 1989; Caldeira, Ciminelli, Dias, & Osseo-Asare, 2003; Ciminelli V. S., 1987; Fan, Markuszewski, & Wheelock, 1984; Koslides & Ciminelli, 1992; Nicol M. J., The Anodic Behaviour of Gold Part II - Oxidation in Alkaline Solutions, 1980; Zhang S. , Oxidation of Refractory Gold Concentrates and Simultaneous Dissolution of Gold in Aerated Alkaline Solutions, 2004). The following sections delve into the reaction chemistries of atmospheric oxidation of iron sulphide ores. While many of the cited works below also contain kinetic data and results, these will not be included here in this work due to the nature of the samples tested from the White Mountain mine. Natural ores – such as those used for this work – have a myriad of different iron oxides, iron sulphides, silicates, clays, native sulphur, and other impurities. As such, without employing some sort of sequential leach procedure, it is impossible to isolate iron sulphides specifically from these samples and test the kinetics of pyrite and marcasite. Future work may use such steps and include test work on pure pyrite and marcasite samples to understand the kinetics of these minerals. For more reading on the kinetics of iron sulphide oxidation, please see the most heavily cited works below. 2.5.1 Acidic Oxidation of Pyrite, Marcasite, and Arsenopyrite  The oxidation and pre-treatment of mineral ores containing iron sulphides such as pyrite, marcasite, and arsenopyrite typically occurs in acidic media, allowing the pH of slurries to decrease until they reach equilibrium. Operators use acids to dissolve ores and oxidants such as oxygen (either pure, or in compressed air) and ferric iron to oxidize iron and sulphur to ferric and sulphate, respectively. The extents of oxidation and dissolution achieved can be quite high, upwards of 90%, as previously discussed. In many cases, autoclaves are used to drive up pressures and temperatures to aid in this oxidation process and decrease residence times for the ore to remain in the pre-treatment stage. The downside to this is that acidic slurries cannot be directly leached with cyanide, as alkaline conditions are required to prevent formation and volatilization of toxic hydrogen cyanide gas (Komnitsas & Pooley, 1989; La Brooy, Linge, & Walker, 1994; Marsden & House, 2006). The process chemistry for acidic oxidation of the three main gold-bearing iron sulphides is shown below (Bonnissel-Gissinger, Alnot, Ehrhardt, & Behra, 1998; Brown & Jurinak, 1989; Chandra & Gerson, 2010; Goldhaber, 1983; Lowson, 1982; Nicol, Miki, Zhang, & Basson, 2013; Sun, Chen, Zou, Shu, & Ruan, 2015; Walker, Schreiber, & Rimstidt, 2006; Zhu, Li, & Wadsworth, Kinetics of 38  the Transpassive Oxidation of Pyrite, 1992; Zhu, Li, & Wadsworth, Characterization of Surface Layers Formed During Pyrite Oxidation, 1994): “Incomplete” oxidation of pyrite/marcasite to ferrous iron and native sulphur Reduction: O2(aq) + 4H+(aq) + 4e- = 2H2O       (eq. 67a) Oxidation: FeS2(s) = Fe+2(aq) + 2S0(s) + 2e-    E° = 0.39 V (SHE) (eq. 67b) Overall Redox: (1)(Reduction) + (2)(Oxidation) 2FeS2(s) + O2(aq) + 4H+(aq) = 2Fe+2(aq) + 4S0(s) + 2H2O(l)     (eq. 67)  “Incomplete” oxidation of pyrite/marcasite to goethite, native sulphur, and acid Reduction: O2(aq) + 4H+(aq) + 4e- = 2H2O(l)      (eq. 68a) Oxidation: FeS2(s) + 2H2O(l) = FeOOH(s) + 2S0(s) + 3H+(aq) + 3e-    (eq. 68b) Overall Redox: (3)(Reduction) + (4)(Oxidation) 4FeS2(s) + 3O2(aq) + 2H2O(l) = 4FeOOH(s) + 8S0(s)      (eq. 68) Oxidation of pyrite/marcasite to ferrous iron, sulphate, and acid Reduction: O2(aq) + 4H+(aq) + 4e- = 2H2O(l)      (eq. 69a) Oxidation: FeS2(s) + 8H2O(l) = Fe+2(aq) + 2SO4-2(aq) + 16H+(aq) + 14e-   (eq. 69b) Overall Redox: (14)(Reduction) + (4)(Oxidation) 2FeS2(s) + 7O2(aq) + 2H2O(l) = 2Fe+2(aq) + 4SO4-2(aq) + 4H+(aq)    (eq. 69) Oxidation of pyrite/marcasite to ferric iron, sulphate, and acid Reduction: O2(aq) + 4H+(aq) + 4e- = 2H2O(l)      (eq. 70a) Oxidation: FeS2(s) + 8H2O(l) = Fe+3(aq) + 2SO4-2(aq) + 16H+(aq) + 15e-   (eq. 70b) Overall Redox: (15)(Reduction) + (4)(Oxidation) 4FeS2(s) + 15O2(aq) + 2H2O(l) = 4Fe+3(aq) + 8SO4-2(aq) + 4H+(aq)    (eq. 70) Oxidation of pyrite/marcasite to ferric hydroxide, sulphate, and acid Reduction: O2(aq) + 4H+(aq) + 4e- = 2H2O(l)      (eq. 71a) Oxidation: FeS2(s) + 11H2O(l) = Fe(OH)3(s) + 2SO4-2(aq) + 19H+(aq) + 15e-   (eq. 71b) Overall Redox: (15)(Reduction) + (4)(Oxidation) 4FeS2(s) + 15O2(aq) + 14H2O(l) = 4Fe(OH)3(s) + 8SO4-2(aq) + 16H+(aq)   (eq. 71) Oxidation of ferrous iron by oxygen to ferric iron Reduction: O2(aq) + 4H+(aq) + 4e- = 2H2O(l)      (eq. 72a) Oxidation: Fe+2(aq) = Fe+3(aq) + e-     E° = -0.70 V (SHE) (eq. 72b) 39  Overall Redox: (1)(Reduction) + (4)(Oxidation) 4Fe+2(aq) + O2(aq) + 4H+(aq) = 4Fe+3(aq) + 2H2O(l)      (eq. 72) Oxidation of pyrite/marcasite by ferric iron to ferrous iron, sulphate, and acid FeS2(s) + 14Fe+3(aq) + 8H2O(l) = 15Fe+2(aq) + 2SO4-2(aq) + 16H+(aq)    (eq. 73) Oxidation of pyrite/marcasite to goethite, sulphate, and acid Reduction: O2(aq) + 4H+(aq) + 4e- = 2H2O(l)      (eq. 74a) Oxidation: FeS2(s) + 10H2O(l) = FeOOH(s) + 2SO4-2(aq) + 19H+(aq) + 15e-   (eq. 74b) Overall Redox: (15)(Reduction) + (4)(Oxidation) 4FeS2(s) + 15O2(aq) + 10H2O(l) = 4FeOOH(s) + 8SO4-2(aq) + 16H+(aq)    (eq. 74) Arsenopyrite dissolution in acidic systems involves the formation of arsenious- and arsenic- acids of varying degrees of protonation. As arsenopyrite is oxidized at a low pH, arsenic acid (H3AsO4) is evolved, with iron oxidized to the ferric (Fe+3) ion, arsenic oxidized to the As (V) species, and sulphur sequentially oxidized to various intermediate thionic compounds (including thiosulphate (S2O3-2) and finally to the sulphate (SO4-2) species. Given below are a few of the key reactions identified by researchers (Asta, Cama, Ayora, Acero, & de Giudici, 2010; Beattie & Poling, 1987; Buckley & Walker, 1988; Corkhill & Vaughan, 2009; Fernandez, Linge, & Wadsley, Oxidation of Arsenopyrite (FeAsS) in Acid Part I: Reactivity of Arsenopyrite, 1996; Fernandez, Linge, & Willing, Oxidation of Arsenopyrite (FeAsS) in Acid Part II: Stoichiometry and Reaction Scheme, 1996; Lengke, Sanpawanitchakit, & Tempel, 2009; Papangelakis & Demopoulos, Acid Pressure Oxidation of Arsenopyrite: Part I, Reaction Chemistry, 1990; Ruitenberg, Hansford, Reuter, & Breed, 1999) (Walker, Schreiber, & Rimstidt, 2006; Yu, Zhu, Gao, Gammons, & Li, 2007): “Incomplete” oxidation of arsenopyrite to ferrous iron, arsenic acid, and native sulphur 2FeAsS(s) + 5O2(aq) + 10H+(aq) = 2Fe+2(aq) + 2H3AsO4(aq) + 2S0(s) + 2H2O(l)   (eq. 75) Or: 4FeAsS(s) + 7O2(aq) + 8H+(aq) + 2H2O(l) = 4Fe+2(aq) + 4H3AsO4(aq) + 4S0(s)   (eq. 76) “Incomplete” oxidation of arsenopyrite by ferric iron to form ferrous iron, arsenic acid, and native sulphur FeAsS(s) + 7Fe+3(aq) + 4H2O(l) = 8Fe+2(aq) + H3AsO4(aq) + S0(s) + 5H+(aq)   (eq. 77) Oxidation of arsenopyrite by ferric iron to form ferrous iron, arsenite (H3AsO3), and sulphate FeAsS(s) + 11Fe+3(aq) + 7H2O(l) = 12Fe+2(aq) + H3AsO3(aq) + SO4-2(aq) + 11H+(aq)  (eq. 78) 40  Oxidation of arsenite to various levels of protonated arsenic acid (reactions in order of occurrence with decreasing pH) 2H3AsO3(aq) + O2(aq) = 2HAsO4-2(aq) + 4H+(aq)      (eq. 79) Or: 2H3AsO3(aq) + O2(aq) = 2H2AsO4-(aq) + 2H+(aq)      (eq. 80) Or: 2H3AsO3(aq) + O2(aq) = 2H3AsO4(aq)       (eq. 81) Oxidation of arsenopyrite by ferric iron and oxygen to form ferrous iron, arsenite, and sulphate 4FeAsS(s) + 11O2(aq) + 6H2O(l) = 4Fe+2(aq) + 4H3AsO3(aq) + 4SO4-2(aq)   (eq. 82) Or: FeAsS(s) + 11Fe+3(aq) + 7H2O(l) = 12Fe+2(aq) + H3AsO3(aq) + SO4-2(aq) + 11H+(aq)  (eq. 83) Oxidation of arsenopyrite to form ferric iron, arsenic acid, and sulphate Step 1: 4FeAsS(s) + 13O2(aq) + 6H2O(l) = 4Fe+2(aq) + 4H3AsO4(aq) + 4SO4-2(aq)  (eq. 84a) Step 2: 4Fe+2(aq) + O2(aq) + 4H+(aq) = 4Fe+3(aq) + 2H2O(l)     (eq. 84b) Overall Reaction: 2FeAsS(s) + 7O2(aq) + 2H2O(l) = 2Fe+3(aq) + 2H3AsO4(aq) + 2SO4-2(aq) (eq. 84) Oxidation of arsenopyrite to ferric hydroxide, arsenite, and sulphate Step 1: 4FeAsS(s) + 11O2(aq) + 6H2O(l) = 4Fe+2(aq) + 4H3AsO3(aq) + 4SO4-2(aq)  (eq. 85a) Step 2: 4Fe+2(aq) + O2(aq) + 10H2O(l) = 4Fe(OH)3(s) + 8H+(aq)    (eq. 85b) Overall Reaction: FeAsS(s) + 3O2(aq) + 4H2O(l) = Fe(OH)3(s) + H3AsO3(aq) + SO4-2(aq) + 2H+(aq)              (eq. 85) Oxidation of arsenopyrite to ferric hydroxide, arsenic acid, and sulphate Reduction: O2(aq) + 4H+(aq) + 4e- = 2H2O(l)      (eq. 86a) Oxidation: FeAsS(s) + 11H2O(l) = Fe(OH)3(s) + HAsO4-2(aq) + SO4-2(aq) + 18H+(aq) + 14e-           E° = -0.569 V  (eq. 86b) Overall Redox: (14)(Reduction) + (4)(Oxidation) 2FeAsS(s) +7O2(aq) + 8H2O(l) = 2Fe(OH)3(s) + 2HAsO4-2(aq) + 2SO4-2(aq) + 8H+(aq)   (eq. 86) Arsenopyrite dissolution releases aqueous arsenic species into the bulk solution; this becomes a problem for industry operators when it comes to treating and storing tails residues and especially mine or process wastewaters. Arsenic and other base metal fixation in the form of precipitation is therefore necessary when processing ores that contain these elements, as will be discussed below. 41  While there exist numerous precipitates that form after oxidation of pyrite and marcasite, only a few environmentally friendly options exist for the capture of aqueous arsenic in industrial applications. Both acidic and aqueous oxidation of arsenopyrite releases aqueous arsenic species such as arsenite or arsenate into the bulk solution, and it is this aqueous arsenic that poses a major health and environmental hazard to the surrounding natural habitat (Demopoulos, 2009; Filippou & Demopoulos, 1997; Fujita, et al., Effect of pH on Scorodite Synthesis by Oxidation of Ferrous Ions: Physical Properties and Stability of the Scorodite, 2009; Fujita, et al., Novel Atmospheric Scorodite Synthesis by Oxidation of Ferrous Sulfate Solution Part 1, 2008; Fujita, Fujieda, Shinoda, & Suzuki, 2012; Langmuir, Mahoney, & Rowson, 2006; Paktunc, Dutrizac, & Gertsman, 2008). In the past, Jarosites have been used to fixate iron, sulphate, and other base metals from solution in order to prevent contamination of the natural environment surrounding mining operations, but these precipitates have been found to have unacceptably high solubilities in ground-waters of varying pH ranges, as well as undesirable settling properties due to their amorphous and gel-like consistencies (Demopoulos, 2009; Marsden & House, 2006; Paktunc, Dutrizac, & Gertsman, 2008; Yannopoulos, 1991). Currently, the precipitate of choice is ferric arsenate, or Scorodite (depending on the degree of hydration that this compound achieves) (Filippou & Demopoulos, 1997; Fujita, et al., Effect of pH on Scorodite Synthesis by Oxidation of Ferrous Ions: Physical Properties and Stability of the Scorodite, 2009; Fujita, et al., Novel Atmospheric Scorodite Synthesis by Oxidation of Ferrous Sulfate Solution Part 1, 2008; Fujita, Fujieda, Shinoda, & Suzuki, 2012; Langmuir, Mahoney, & Rowson, 2006). Various authors, including Dutrizac, Demopoulos, Filipou, and especially Fujita et al. have all hypothesized the mechanism of Scorodite precipitation. The general reactions for precipitation of Scorodite is given below (Demopoulos, 2009; Filippou & Demopoulos, 1997; Fujita, et al., Effect of pH on Scorodite Synthesis by Oxidation of Ferrous Ions: Physical Properties and Stability of the Scorodite, 2009; Fujita, et al., Novel Atmospheric Scorodite Synthesis by Oxidation of Ferrous Sulfate Solution Part 1, 2008): Precipitation of ferric arsenate or Scorodite from aqueous arsenic acid and ferric sulphate 2H3AsO4(aq) + Fe2(SO4)3(aq) + 4H2O(l) = 2FeAsO4·2H2O(s) + 3H2SO4(aq)   (eq. 87) The problem of base metal fixation (and especially arsenic fixation) has been the subject of intense scrutiny over at least the past three decades and is becoming increasingly important as environmental regulations surrounding precious metals mining operations are tightened. Scorodite has thus far proven to be extremely stable in circum-neutral pH groundwater, able to form large, crystalline, fast-settling precipitates which aid in dewatering mining tails, and above all, able to fixate arsenic and other base metals from solution so that contamination of the environment is avoided (Asta, et al., 2013; Demopoulos, 2009; Filippou & Demopoulos, 1997; Fujita, et al., Effect of pH on Scorodite Synthesis by Oxidation of Ferrous Ions: 42  Physical Properties and Stability of the Scorodite, 2009; Fujita, Fujieda, Shinoda, & Suzuki, 2012; Langmuir, Mahoney, & Rowson, 2006; Paktunc, Dutrizac, & Gertsman, 2008) 2.5.2 Alkaline Oxidation of Pyrite, Marcasite, and Arsenopyrite  Ciminelli studied the oxidation of iron sulphide minerals in alkaline solutions containing sodium carbonate or sodium hydroxide, both with and without the addition of cyanide to study the dissolution and solubility of these minerals, the formation of various iron-cyanide complexes and arsenic- and sulphur-containing compounds, and their effect on precious metals recovery from natural ores (Ciminelli V. S., 1987). Other oft-cited authors, such as Demopoulos, Hiskey, Papangelakis, and Wadsworth, have all contributed to the subject of alkaline oxidation of iron sulphide minerals, especially pyrite, marcasite and arsenopyrite. These studies on the reaction chemistries, observed kinetics, and by-products precipitated after oxidation have built a general understanding of possible pre-treatment processes that can be used in industry to enhance precious metals recovery. Many of the oxidation reactions that occur for iron sulphides are the same as those described above for acidic oxidation, but with the addition of the formation of water because of the reaction between H+ ions and OH- ions. In practice, the acidic species (H+) would be immediately consumed by excess OH- in solution, and so water is formed. This can be seen below in the following equation: H+(aq) + OH-(aq) = H2O(l)         (eq. 87) For clarity, a few key equations are given below (Ahlberg, Forssberg, & Wang, 1990; Caldeira, Ciminelli, Dias, & Osseo-Asare, 2003; Ciminelli & Osseo-Akare, Kinetics of Pyrite Oxidation in Sodium Carbonate Solutions, 1995; Ciminelli & Osseo-Akare, Kinetics of Pyrite Oxidation in Sodium Hydroxide Solutions, 1995; Ciminelli V. S., 1987; Goldhaber, 1983; Nicholson, Gillham, & Reardon, Pyrite Oxidation in Carbonate-Buffered Solution: 1. Experimental Kinetics, 1988; Tao, Li, Richardson, & Yoon, 1994): “Incomplete” oxidation of pyrite/marcasite, leaving native sulphur as a by-product Reduction: O2(aq) + 2H2O(l) + 4e- = 4OH-(aq)      (eq. 88a) Oxidation: FeS2(s) + 3OH-(aq) = Fe(OH)3(s) + S0(s) + 3e-     (eq. 88b) (3)(Reduction) + (4)(Oxidation) Overall Reaction: 4FeS2(s) + 3O2(aq) + 6H2O(l) = 4Fe(OH)3(s) + 8S0(s)   (eq. 88) Or: 4FeS2(s) + 3O2(aq) + 2H2O(l) = 4FeOOH(s) + 8S0(s)      (eq. 89) Oxidation of pyrite/marcasite by ferric iron to ferrous iron, sulphate (and acid, which is immediately neutralized in alkaline environments) 43  FeS2(s) + 14Fe+3(aq) + 8H2O(l) = 15Fe+2(aq) + 2SO4-2(aq) + 16H+(aq)    (eq. 90a) Aside: H+(aq) + OH-(aq) = H2O(l)        (eq. 90b) Overall Reaction: (1)(a) + (16)( b) FeS2(s) + 14Fe+3(aq) + 16OH-(aq) = 15Fe+2(aq) + 2SO4-2(aq) + 8H2O(l)    (eq. 90) Oxidation of pyrite/marcasite to ferrihydrite and sulphate 4FeS2(s) + 15O2(aq) + 16OH-(aq) = 4Fe(OH)3(s) + 8SO4-2(aq) + 2H2O(l)   (eq. 91) Alternatively, oxidation of pyrite/marcasite to goethite and sulphate 4FeS2(s) + 15O2(aq) + 16OH-(aq) = 4FeOOH(s) + 8SO4-2(aq) + 6H2O(l)    (eq. 92) Oxidation of pyrite/marcasite to stable hematite and sulphate 4FeS2(s) + 15O2(aq) + 16OH-(aq) = 2Fe2O3(s) + 8SO4-2(aq) + 8H2O(l)    (eq. 93) This last reaction produces hematite, which is the most stable iron precipitate, but only occurs after goethite loses water at higher temperatures, which is possibly why this reaction occurs in pressure oxidation of pyrite, rather than under atmospheric pressures. Goethite may revert to hematite during the water recovery process if oxidized and leached residues (or “tails”) undergo a drying process, or if enough time has passed where the structure of the precipitate surface layer changes in oxidative leach conditions (Ahlberg, Forssberg, & Wang, 1990; Caldeira, Ciminelli, Dias, & Osseo-Asare, 2003). However, the timescale in which natural reversion of goethite to hematite may be in the order of thousands, or hundreds of thousands of years, as evidenced by natural goethite deposits found around the world in mineral deposits containing iron oxy-hydroxides (Dreisinger, 2014). The dissolution of arsenopyrite in alkaline solutions is, again, very similar to that of pyrite and marcasite. The presence of aqueous arsenic in the system, typically oxidized by either oxygen or ferric iron, differentiates the reaction chemistry of the system. Operators must prevent contamination of wastewater effluents with arsenic and other base metals that may be released by oxidative pre-treatment. As pH of the system increases, arsenic acid (H3AsO4) releases each of its protons (H+) in succession as pH increases. Finally, only the AsO4-3 species is left in solution at high pH (>~12). A few authors have studied this system, and as in the section on aqueous pressure oxidation, the dissolution reactions are given below (Amoah-Forson, 1986; Asta, et al., 2013; Asta, Cama, Ayora, Acero, & de Giudici, 2010; Bhakta, Langhans, & Lei, 1989; Buckley & Walker, 1988; Corkhill & Vaughan, 2009; Koslides & Ciminelli, 1992; Nicol & Guresin, 2003; Ruitenberg, Hansford, Reuter, & Breed, 1999; Yu, Zhu, Gao, Gammons, & Li, 2007): Oxidation of arsenopyrite to hematite, sulphate, and arsenate (AsO4-3) Step 1: 2FeAsS(s) + 7O2(aq) + 4OH-(aq) = 2Fe+3(aq) + 2AsO4-3(aq) + 2SO4-2(aq) + 2H2O(l) (eq. 94a) 44  Step 2: 2Fe+3(aq) + 6OH-(aq) = Fe2O3(s) + 3H2O(l)      (eq. 94b) Overall Reaction: 2FeAsS(s) + 7O2(aq) + 10OH-(aq) = Fe2O3(s) + 2AsO4-3(aq) + 2SO4-2(aq) + 5H2O(l)             (eq. 94) Oxidation of arsenopyrite to ferric hydroxide, arsenate, and sulphate Reduction: O2(aq) + 2H2O(l) + 4e- = 4OH-(aq)      (eq. 95a) Oxidation: FeAsS(s) + 19OH-(aq) = Fe(OH)3(aq) + AsO4-3(aq) + SO4-2(aq) + 8H2O(l) + 14e-  (eq. 95b) Overall Redox: (14)(Reduction) + (4)(Oxidation) Overall Reaction: FeAsS(s) + 15OH-(aq) + O2(aq) = Fe(OH)3(s) + AsO4-3(aq) + SO4-2(aq) + 6H2O(l)             (eq. 95) Oxidation of arsenopyrite to ferric hydroxide, arsenic acid (HAsO4-2), and sulphate Reduction: O2(aq) + 2H2O(l) + 4e- = 4OH-(aq)      (eq. 96a) Oxidation: FeAsS(s) + 11H2O(l) = Fe(OH)3(s) + HAsO4-2(aq) + SO4-2(aq) + 18H+(aq) + 14e- (eq. 96b) Overall Redox: (14)(Reduction) + (4)(Oxidation) 4FeAsS(s) + 72H2O(l) + 14O2(aq) = 4Fe(OH)3(aq) + 4HAsO4-2(aq) + 4SO4-2(aq) + 72H+(aq) + 56OH-(aq)             (eq. 96c) Aside: H+(aq) + OH-(aq) = H2O(l)        (eq. 96d) Overall Reaction: (1)(c) + (16)(d) 2FeAsS(s) + 7O2(aq) + 8OH-(aq) = 2Fe(OH)3(aq) + 2HAsO4-2(aq) + 2SO4-2(aq)   (eq. 96) Oxidation of arsenopyrite to ferric hydroxide (or hydrous ferric oxide, HFO), arsenite, and sulphate FeAsS(s) + 3O2(aq) + 4H2O(l) = Fe(OH)3(s) + H3AsO3(aq) + SO4-2(aq) + 2H+(aq)   (eq. 97a) H+(aq) + OH-(aq) = H2O(l)         (eq. 97b) Overall Reaction: (1)(a) + (2)(b) FeAsS(s) + 3O2(aq) + 2OH-(aq) + 2H2O(l) = Fe(OH)3(s) + H3AsO3(aq) + SO4-2(aq)  (eq. 97) Arsenite is also oxidized in solution by both oxygen and ferric iron to produce arsenate, as shown below: Oxidation of arsenite to various forms of arsenic acid 2H3AsO3(aq) + O2(aq) = 2HAsO4-2(aq) + 4H+(aq)      (eq. 98a) H+(aq) + OH-(aq) = H2O(l)         (eq. 98b) Overall Reaction: (1)(a) + (4)(b) 2H3AsO3(aq) + O2(aq) + 4OH-(aq) = 2HAsO4-2(aq) + 4H2O(l)     (eq. 98) Alternatively: Reduction: O2(aq) + 2H2O(l) + 4e- = 4OH-(aq)      (eq. 99a) 45  Oxidation: AsO3-3(aq) + 2OH-(aq) = AsO4-3(aq) + H2O(l) + 2e-    (eq. 99b) Overall Redox: (1)(Reduction) + (2)(Oxidation) 2AsO3-3(aq) + O2(aq) = 2AsO4-3(aq)        (eq. 99) The rate of arsenopyrite oxidation is generally seen to be higher than that of pyrite and marcasite, possibility due to the comparative instability of the crystal structure of arsenopyrite caused by displacement of atoms surrounding the As-S dianion pair (Asta, Cama, Ayora, Acero, & de Giudici, 2010; Komnitsas & Pooley, 1989; Tossel, Vaughan, & Burdett, 1981; Walker, Schreiber, & Rimstidt, 2006).  As will be shown below, tested mineral ores from the White Mountain mine contain mainly pyrite and marcasite as gold-bearing minerals of interest, and very little arsenopyrite was present in these ores. This work was originally undertaken with the goals described in Chapter 1. They are given below: 1. To record rates of oxidation of ore samples obtained from the White Mountain mine and relating this back to the varying contents of pyrite, marcasite, arsenopyrite, silicate material, as well as any other significant gold-containing minerals in each ore. Specifically: a. Measure the rates of reaction and develop models to confirm the mode of oxidation b. Determine the controlling mechanism of the reaction and distinguish between factors that control the rate of reaction (e.g. pyrite and marcasite oxidation conforming to various leaching models such as mass-transfer through ion-diffusion, or chemical reaction control) c. Determine the activation energies for iron sulphide minerals in each sample of ore that is tested 2. To delineate the effects of alkaline oxidative pre-treatment on the extraction of gold from each natural ore sample by cyanidation. Parameters investigated include caustic dosage, air sparging levels, and temperature. 3. To describe the oxidation products that resulted from the alkaline pre-treatment step, and any influence they may have on the overall extraction of gold The experimental work undertaken to accomplish these goals is described below in subsequent chapters, as well as the results and discussion of findings obtained.46  Chapter 3: Experimental Work  The experiments undertaken for this work consisted of a conventional gold leaching step (“cyanidation”) for all experimental trials, with an alkaline atmospheric oxidation pre-treatment step conducted only for certain trials, varying, one at a time, different physical variables such as temperature and rate of air input under constant pressure to delineate the effect of each variable on overall gold extraction.  3.1 Materials 3.1.1 Reagents  All reagents used in this work were either of reagent, or analytical grade and were used without further purification. Except for the comminution of the natural ore samples, where process or tap water was used, only deionized water or ultrapure (18 MΩ) filtered water was used. For pre-oxidative treatment prior to leaching tests, breathing grade (99.99% pure) compressed air was supplied by Praxair. Given below in Table 6 is a compiled list of all chemical reagents used, their respective purities, forms, and sources. Table 6: Chemical Reagents Used for Leaching Tests Reagent Concentration Grade/Purity Source Form Sodium Hydroxide (NaOH)  98% (ACS Reagent Grade) Anachemia Solid (Pellets) Sodium Hydroxide (NaOH)  98% (ACS Reagent Grade) Alfa Aesar/Thermo Fisher Scientific Solid (Pellets) Sodium Hydroxide (NaOH) 1.0 N (ACS Reagent Grade) BDH Chemicals/VWR Analytical Solution Sodium Cyanide (NaCN)  95% (ACS Reagent Grade) Anachemia (VWR) Solid (Powder) Silver Nitrate (AgNO3) 0.1 N Certified Analytical Grade Fisher Scientific Solution Potassium Iodide (KI)  (ACS Reagent Grade) ACROS Solid (Powder) Calcium Hydroxide (Ca(OH)2)  95% (ACS Reagent Grade) Anachemia (VWR) Solid (Powder)  47  3.1.2 Ore Samples  Given below in Table 7 is an elemental summary of each ore provided by the Eldorado Gold Corporation. These ores were sourced from the White Mountain mine in China. Table 7: Elemental Summary of All White Mountain Ore Samples Sample #: 165 166 167 168 169 200 201 Au (g/mt) 14.16 24.28 19.35 19.14 7.57 4.83 4.11 Ag (g/mt) 9.50 11.40 9.90 7.30 8.55 1.60 1.35 Fe (%) 1.79 5.32 8.87 9.92 11.35 4.02 4.33 S (total, LECO) (%) 0.88 5.18 7.94 10.46 13.35 4.07 3.89 Ca (%) 0.19 0.08 0.07 0.11 0.07 0.05 0.15 P (%) 0.04 0.09 0.11 0.13 0.06 0.05 0.05 Mg (%) 0.12 0.09 0.07 0.08 0.07 0.08 0.13 Ti (%) 0.03 0.04 0.04 0.04 0.04 0.06 0.05 Al (%) 1.00 1.40 1.42 1.60 1.22 1.65 1.83 Na (%) 0.07 0.02 0.01 0.01 0.01 0.02 0.01 K (%) 0.28 0.42 0.37 0.42 0.41 0.44 0.47 Ba (ppm) 20188.00 557.00 492.00 443.00 141.50 2989.00 6751.50 As (ppm) 172.00 470.00 535.00 769.00 1124.00 389.00 634.00 Cr (ppm) 202.00 174.00 173.00 140.00 194.00 202.00 181.50 Cu (ppm) 111.80 47.50 30.40 20.20 20.15 25.80 14.80 Pb (ppm) 89.10 51.90 72.50 81.80 6.20 8.60 9.90 Zn (ppm) 57.00 31.00 23.00 24.00 20.50 29.00 27.50 Ni (ppm) 18.90 43.00 58.30 85.70 69.65 25.60 45.70 48  Sample #: 165 166 167 168 169 200 201 Hg (ppm) 10.64 30.20 33.07 46.42 70.25 24.57 32.59  Given below in Table 8 are results from Quantitative X-Ray Diffraction (QXRD) analysis for select ores chosen for experimentation, performed by the Earth and Ocean Sciences department at UBC. Table 8: Summary of Results from XRD Analysis and Rietveld Refinement (QXRD) Mineral Ideal Formula #165 167 169 201 Actinolite Ca2(Mg,Fe2+)5Si8O22(OH)2    0.3 Barite BaSO4 6.2 1.8  1.7 Goyazite SrAl3(PO4) 2(OH)5·H2O  0.7 0.3 0.3 Gypsum CaSO4·2H2O  2.4 3.3 2.8 Hematite -Fe2O3 0.3    Illite/Muscovite 2M1 K0.65Al2.0Al0.65Si3.35O10(OH)2 / KAl2AlSi3O10(OH)2 1.3 2.7 4.3 2.9 Iron Fe 0.4    Jarosite K2Fe63+(SO4)4(OH)12   0.5 0.7 Kaolinite Al2Si2O5(OH)4 0.8 1.8 1.5 1.8 Magnetite Fe3O4 0.1    Marcasite FeS2  4.5 12.7 0.7 Melanterite FeSO4·7H2O  2.5   Pyrite FeS2  5.6 11.0 3.1 Quartz SiO2 90.9 78.0 65.3 85.7 Sulfur S   1.1  Total  100.0 100.0 100.0 100.0 49  For the table above, the presence of native iron (Fe) may be attributed to the comminution process where iron shavings from the crushing and grinding equipment may have been included to sample 165, which was the first sample to undergo this process. Native sulphur may be naturally occurring in sample 169; however, it is also likely that the detection of sulphur in this sample may be due to the wet grinding process, which allowed some sulphides in the ore to be partially oxidized.  XRD diffraction patterns for all untreated (“Head”) samples in this work are provided in Appendix C. The diffraction patterns for leached residues (“Tails”) used to analyze phases present after oxidation and cyanidation tests have not been included. The head samples were analyzed with Rietveld Refinement performed by the Earth and Ocean Sciences department at the UBC Point Grey (Vancouver) campus. Other diffraction pattern analyses were performed by the author, using the “Match!” software by Dr. Holger Pulz (Crystal Impact, 2018). Given below is an example of the diffraction pattern for untreated sample 165.  Figure 18: XRD Pattern for untreated ("Head") sample 165; analysis and Rietveld Refinement performed by the department of Earth and Ocean Sciences at UBC Note also that the names given to the ores received from the White Mountain mine were assigned prior to shipment, and these names do not necessarily reflect the actual compositions of each ore. From internal documents and communication with the sponsor, it was understood that the ores were mainly comprised of 2Th Degrees8075706560555045403530252015105Sqrt(Counts)220200180160140120100806040200-201AR_165.raw_1 Quartz low 90.86 %Barite 6.18 %Kaolinite 1A 0.80 %Iron-alpha 0.40 %Magnetite 0.12 %Illite/Muscovite 2M1 1.28 %Hematite 0.35 %50  Quartz and other silicates, with the bulk of iron and sulphur associated with pyrite, marcasite, and potentially some arsenopyrite (ALS Ammtec Sydney, 2012; Dreisinger, 2014). It is also worthwhile to note the high quartz content present in each ore, as well as the closely matching numbers for the iron and sulphur contents of each ore. This, along with the results of the Quantitative XRD analysis, provide evidence that all iron in the ores are associated with the sulphur as iron sulphides such as pyrite or marcasite, and any extra iron is likely found as iron oxides or -hydroxides. Any extra sulphur may be present as native sulphur or as ferrous iron sulphate hydrate, which may have been formed from iron sulphides that were partially oxidized from previous exposure to the natural environment and the wet grinding process. Also important are the low numbers for arsenic content, indicating that very little arsenopyrite is present in each ore. This is confirmed using QXRD testing. A few notable results can be inferred from XRD analysis and subsequent Rietveld analysis. For all samples, quartz is the number one phase present, Other silicate phases include illite/muscovite, and kaolinite. Generally, pyrite and marcasite are the iron sulphide phases present in the ores. Barite is seen in significant quantities for Sample 165. The main iron-bearing phases for this sample are native iron (Fe), and iron-oxides such as hematite (Fe2O3) and magnetite (Fe3O4). Gypsum, a by-product of reaction between slaked lime (Ca(OH)2), water, and sulphuric acid (presumably produced during the grinding stages), is present in significant quantities also. A phase that may be of interest is melanterite (FeSO4·7H2O), a hydrated form of ferrous sulphate. This phase was found to be present in Sample 167 and may be a result of partially oxidized iron sulphides releasing aqueous iron and sulphate during the wet grinding stage. Given below in Figure 19 are dried samples of each ore after crushing and grinding.  Figure 19: Dried ore samples after crushing and grinding 51  After performing the initial tests described above, it became apparent that the similarity of the ores described above allowed for elimination of some samples from future leaching tests. Sample 165 was chosen because of its distinctiveness from the rest of the ores, and Sample 201 was chosen from the two newer samples received from the White Mountain mine, coming from deposits that contained higher concentrations of “refractory” iron-sulphides. Samples 167 and 169 were partially chosen for the high gold and high iron and sulphur contents found in each ore. The similarity in composition of ores 166 – 169 meant that these ores could be interchangeable. For tests described below in Chapter 4, leaching tests were performed on the following ores: Sample 165, 167, 169, and 201.  3.2 Preparation of Ore Samples  To prepare each selected ore sample for leaching, it was first necessary to crush and grind the sample to roughly mimic the conditions used in the White Mountain mine. Through discussion with the industry sponsor, project supervisor, as well as through internal documents, it became apparent that the operator used a P80 grind size of 45μm. This means that roughly 80% of all ground material has a diameter of 45μm or smaller. However, commercial operators generally avoid finer grinding, as increased grind times for ores may lead to higher energy costs. Additionally, over-grinding the ore samples would result in material with a smaller P80 than is possible to achieve at the White Mountain mine. Therefore, a grind size (P80) of 50μm was chosen. P80 grind sizes of all ground ores were determined through laser size analysis (LSA) using a Malvern Mastersizer unit, as well as dry sieving. Given below in Table 9 and Table 10 are the results of the grind tests and chosen grind times necessary to each ore sample down to 50 ± 10μm. Table 9: Grind Calibration Data Obtained Through Sieving and Laser Size Analysis  Sieve Data P80 (μm) LSA Data P80 (μm) Average Data Grind Time Ore 20 40 60 20 40 60 20 40 60 Min. 165 212.00   47.50 163.53 53.12 51.97 187.76 53.12 49.73 52 166      50.00 78.02 59.54 46.10 78.02 59.54 48.05 56 167  115.00   42.50 100.45 39.74 37.19 107.73 39.74 39.85 40 168  110.00 50.00 47.50 103.75 66.00 43.25 106.88 58.00 45.37 51 169      50.00 81.23 53.26 39.44 81.23 53.26 44.72 47 200       72.15 41.40 26.71 72.15 41.40 26.71 30 201       73.09 54.70 37.10 73.09 54.70 37.10 39 52   Table 10: Conservative Grind Times Used to Obtain a P80 Size of 50 μm Ore Type Grind Time (minutes) 165 50 167 35 169 40 201 35  Given below in Figure 20 is a plot of measured P80 values that were used to obtain the calibrated grind times given in the table above.  Figure 20: Plot of measured P80 values obtained using the sieving and Laser Size Analysis techniques  3.3 Experimental Set-up  Leaching experiments conducted for this work were performed using 2000 mL baffled Pyrex reactors from CanSci Glass Products Ltd. A clear plastic lid was fashioned for each reactor to minimize evaporation and chemical volatilization. Temperature control was maintained using a plastic tub as a water bath. The bath 2040608010012014016018020010 20 30 40 50 60 70Measured P80(μm)Grind Time (minutes)16516616716816920020153  had a heating element placed inside it connected to an Omega CN9000A temperature control module which used a thermocouple thermometer to measure the temperature of the water to within 0.1°C. To circulate water within the tub, a Cole-Parmer Model 50006-01 Compact Mixer was run at 1000 RPM to mix water from the opposite side of the tub. An additional thermometer was used to check that water temperature was uniform. Given below in Figure 21 and Figure 22 are schematics of the experimental set-up:  Figure 21: Schematic of the leaching reactor used, with lid attached, and mixing unit suspended above. The holes drilled into the lid were ordered as follows: 1) Air sparging inlet; 2) Dissolved Oxygen electrode; 3) Air vent (if needed); 4, 5, 6) Eh-pH probes and sampling outlet; 7) Impeller shaft inlet  Figure 22: Schematic of the layout of the experimental set-up (overhead view). Components are numbered: 1) Reactor; 2) Heating element; 3) Thermometer; 4) Overhead mixer unit; 5) Humidifier 54  To mix the slurries inside the Pyrex reactors, the same model of mixer was used. During oxidation, air was pumped using air dispersion tubes which are glass tubes with porous ends (glass frits) to ensure proper bubble formation that will sufficiently aerate the slurry. Compressed air was delivered using a two-stage high pressure air regulator which limited air pressure to 10 psi. A Cole-Parmer T-34500-22 flowmeter was then used to regulate the volume of air pumped into the slurry, up to 2.5 L of air per minute. To prevent excessive moisture loss during oxidation (due to air sparging), air was first pumped into sealed Erlenmeyer flasks filled with pure de-ionized water (DI) submerged in the same water bath as the leaching reactors. The humidified air was then forced out of the Erlenmeyer flasks and into the reactors to aerate the slurries. Given below in Figure 23 is a schematic of the flasks that were used as humidifiers.  Figure 23: Schematic of a sealed Erlenmeyer flask, filled with De-Ionized (DI) water, used as a humidifier. Components are numbered as follows: 1) Air input hose; 2) Air dispersion tube and glass frit; 3) Humidified air outlet hose  3.4 Experimental Methods  Leaching tests conducted for this work had three phases: oxidation, pH conditioning, and cyanidation. Baseline cyanidation tests only involved conditioning of the ore slurries to high pH levels, and then a cyanidation phase. Descriptions of analytical methods used to test samples are given below in Appendix B. 55  3.4.1 Oxidation Tests  The method of oxidative pre-treatment experiment in this work was a constant-pH or “high alkalinity” test. This type of experiment involved an initial dosage of sodium hydroxide to leaching reactors, but also a constant maintenance of alkaline conditions in the leaching reactor by monitoring pH and constant dosing additional sodium hydroxide. Caustic sodium hydroxide solution was dosed as necessary to maintain the desired pH level throughout each trial. Alkaline pre-treatment experiments used the same baseline values for physical conditions involved in each trial. These baseline parameters are given below in Table 11: Table 11: Physical Parameters Used for Baseline Oxidation Tests Parameter Value Notes Temperature 35°C Baseline value only Constant throughout oxidation phase Stirring Rate 1250 RPM* Consistent for all oxidation tests Constant throughout oxidation phase Air Input Rate 2.0 L/min at 0.68 atm (10 psi)** Baseline value only Constant throughout oxidation phase Sodium Hydroxide Dosage 7.5 kg/t ore Baseline Value only Initial Dosage (Caustic solution added as needed to maintain pH 11) Pulp Density (w/w) 20% Consistent for all tests Initial Condition Oxidation Time 8 hours Consistent for all oxidation tests Matching times used in the White Mountain Mine * to ensure full agitation, even at the surface of the slurry ** air pressure measured at the air tank regulator; rate of airflow measured at a flowmeter prior to air entering the air moisturizers (see Figure 23) The parameters for the oxidation phase of the tests and their selected values are given below in Table 12:   56  Table 12: Parameters Varied for Oxidation Tests Parameter Tested Values Selected Notes Sodium Hydroxide Dosage 7.5 kg/t 7.5 kg/t initial dosage for alkaline tests, 1.0 M NaOH dosed as necessary to maintain pH 11 Air Input ** 1.5 L/min 2.0 L/min 2.5 L/min For alkaline oxidation studying effect of air input on gold extraction Temperature 35°C 45°C 55°C For alkaline oxidation studying temperature ** air pressure measured at the air tank regulator; rate of airflow measured at a flowmeter prior to air entering the air moisturizers (see Figure 23) To begin each oxidation test, the requisite weighed amounts of de-ionized water, caustic sodium hydroxide (in pellet form), and ground ore were added to the leaching reactor, which was weighed before and after addition of reagents. The reactor was placed into the heated water bath, and had a plastic lid fastened. Next, the mixing unit was started and set at the required stirring rate; an air dispersion tube was lowered into the slurry bubbling air. After this initial set up, physical parameters were measured. These are given below in Table 13, along with the equipment used to measure them and any additional notes: Table 13: Physical Parameters Measured During Oxidation and Cyanidation Tests Parameter Equipment Used Notes pH Oakton pH 150 Multifunction Meter and Probe Calibrated before each test Eh (ORP) Oakton pH 150 Multifunction Meter and Probe Calibrated before each test Dissolved Oxygen (D.O.) Extech Instruments 407510A Dissolved Oxygen Meter and Probe Calibrated before each test Temperature Oakton pH 150 Multifunction Meter and Probe Checked for accuracy before each test  57  At set times during the test, these parameters would be checked. These times were: 0 hours (initial value), 1 hour, 2 hours, 4 hours, 8 hours (end of oxidation test using air sparging), and 24 hours. After recording data for each parameter, the test would be paused to take a sample of leachate solution. This involved shutting off the stirring unit, removing the air dispersion tube from the slurry, and allowing the slurry to settle for at five minutes, until the leachate at the top of the slurry was clear. A 10-mL aliquot of leachate was removed from the reactor and accurately weighed. The test was then allowed to continue after stirring and air input was restarted, and the lid replaced on the leaching reactor. The sample would then be filtered through an IC Millex®-LG 0.20 μm PTFE syringe filter weighed before and after filtering the solution sample, and then weighed again after drying to ensure no fine particulates had been captured by the syringe filter during this process. This solution sample would then be collected with the other samples taken and sent to an external company, AuTec Innovative Extractive Solutions, for analysis for sulphate (SO4-2), as well as other elements in solution. At the eight-hour mark, which was the end of the oxidation phase of each leaching test, after recording the usual physical data, a slurry sample would be taken from the reactor and filtered using a vacuum filtration unit and a Whatman 42 paper filter (porosity diameter of 2 μm). Again, 10 mL of the solution from this filtration would be recovered and used as the eight-hour solution sample. The excess solution would be returned to the reactor. The solid residue recovered from the filtration was sent to Inspectorate International Ltd. (Bureau Veritas S. A.) for total sulphur analysis (using a LECO brand combustion IR detector instrument). After sampling, the reactor was removed from the water bath, dried, and weighed to determine the weight change in the slurry. If the oxidation test was performed at higher temperatures than the baseline value, the water bath would then be reset at 35°C to get ready for the next stages of the leaching test. After weighing, the reactor was replaced in the water bath, with the lid replaced and again tightly sealed. The reactor was left overnight without air or stirring to resume the next phases of the leaching test at the start of the next morning. At the 24-hour mark, the stirring unit would be restarted, and physical data (Eh, pH, D.O., temp.) would again be recorded. Another slurry sample would also be taken with both solids and 10 mL of liquid recovered for analysis. This extra sampling was taken to determine whether any additional oxidation may have occurred overnight even though the oxidation phase was already completed at the eight-hour mark of the test. After the sampling procedure, the reactor was again removed from the water bath, dried, and weighed to get ready for the next stage of the leaching test.  58  3.4.2 pH Conditioning  After the oxidation stage described above, the oxidized ore slurry would often have a lower pH value than was originally recorded. Since cyanidation can only be safely performed at a high pH, an intermediate conditioning phase was undertaken to raise the level of alkalinity in the slurry. Sodium hydroxide was dosed as necessary until a pH of 11 was achieved and the pH of the slurry was seen to be stable. The amount of caustic dosed to each slurry was recorded, and the slurry was then allowed to condition while being stirred vigorously. This process generally took anywhere between 45 to 60 minutes. The pH of the slurry was deemed stable when it did not drop in value by more than 0.05 over a 20-minute period. In some cases, extra caustic sodium hydroxide and conditioning time was required to safely proceed. 3.4.3 Cyanidation Tests  The final stage of the leaching tests performed for this work was cyanidation. In the case of “baseline” cyanidation, only the pH conditioning stage and the gold leaching (cyanidation) stage was performed. After checking that the pH of the slurry was approximately 11, the overhead stirring unit was stopped, and the reactor removed from the water bath, and the outside of the reactor was wiped off to dry it. The requisite dosage of sodium cyanide was added to the reactor and the weight of the reactor was taken. The reactor was then returned to the water bath (already set at 35°C) and the stirring unit was restarted. For the cyanidation phase, no lid was used to achieve better air access to the leach slurry, since no air sparging was being performed in this stage. A compiled list of physical parameters is given below in Table 14: Table 14: Physical Parameters Used for Cyanidation Tests Parameter Value Notes Temperature 35°C Consistent for all cyanidation tests Constant throughout cyanidation phase Stir Rate 2000 RPM* Consistent for all cyanidation tests Constant throughout cyanidation phase Air Input Rate N/A Air naturally mixed into slurry through vigorous agitation (stirring) Sodium Hydroxide Dosage N/A Dosage given in the pH conditioning stage to increase pH of the slurry to 11 Dependent on each type of ore * to ensure full agitation, even at the surface of the slurry, without forced air input 59  At the start of the cyanidation phase, the same physical data were recorded as in the oxidation phase (Eh or ORP, pH, dissolved oxygen (D.O.), and temperature of the slurry). As before, leaching continued with periodic recordings of these parameters, followed by sampling of the leachate. The points of data recording were again 0 hours (initial values), 1 hour, 2 hours, 4 hours, 8 hours, and 24 hours (end of leaching tests). The leachate sampling was again performed by stopping the stirring unit and allowing solids to settle for 5 – 10 minutes before removing a 10-mL aliquot of clear solution which was weighed. Stirring was restarted immediately after sampling, and leaching could then continue. At hour 8, after recording physical data and sampling, the leaching reactor was removed from the water bath briefly, dried and then weighed. The reactor was then replaced into the water bath and stirring was again started. The reactor would then be left overnight to continue the leaching phase until the 24-hour mark (the next morning). At the start of the next day, at approximately the 23-hour mark, the reactor would again be removed and weighed, and any weight loss associated with evaporation would be compensated by addition of de-ionized water. The leaching reactor would then be returned to the water bath and the leaching test would allow to go until completion at 24 hours. At the end of the test, after recording physical data, the overhead stirring units would be stopped, and the leaching reactors would be removed from the water bath. The solids in the leached slurry would then be allowed to settle for 30 to 45 minutes, before being filtered through a Whatman 42 paper filter (again using a vacuum filtration unit). The volume and weight of the pregnant leach solution (PLS) would be accurately determined using a volumetric measuring cylinder. A final 24-hour aliquot of solution would be taken from this solution. The leached residue (“tails”) would be washed using de-ionized water roughly equivalent to two bed-volumes of the leaching residue (at least 200 mL). After filtration, the volume and weight of the wash water would be accurately determined and recorded, and a final 10-mL sample would be taken from this solution. The 1-, 2-, 4-, 8-, 24-hour, and wash solution samples would then be sent to Inspectorate for aqueous gold analysis, by Atomic Adsorption Spectrophotometry (AAS). The leaching residue, or tails, would then be dried overnight in an 80°C oven. The filter cake would be rolled to separate solids into powder form again and weighed accurately. The filter paper, and any paper towels used to store the filter cake in the oven would be weighed to determine any possible solids loss. A representative sample would be removed and sent to Inspectorate for gold fire assay (using ICP-AES/OES) as well as total sulphur analysis (LECO analysis).60  Chapter 4: Results and Discussion  This chapter summarizes the results of oxidation and cyanidation tests, as well as XRD phase analysis performed on leached residues (“tails”). All tests were performed using the experimental setup outlined in Chapter 3 above. Baseline cyanidation was performed to determine the level of gold extraction prior to or without any pre-oxidative treatment. Alkaline pre-oxidation and subsequent cyanidation was performed on the four ores selected for testing, with three variables considered: caustic consumption, rate of air input, and oxidation temperature. Sample calculations for mass balance tables are given in Appendix C. Example mass balance tables for tests with and without alkaline pre-treatment are given in Appendix D.   4.1 Results of Baseline Cyanidation Tests  All cyanidation, including the baseline cyanidation tests without oxidative pre-treatment, were performed according to the experimental setup given in Chapter 3. Briefly, the specifics for the following tests are: 35°C cyanidation temperature, 2000 RPM stirring rate, no air sparging, and no initial dosage of sodium hydroxide. For each ore sample, baseline cyanidation was conducted in duplicate and the average between the two tests were taken to be the “average baseline cyanidation” level that each ore would yield relative to their own assayed head grades. These levels were compared with the gold extractions resulting after alkaline oxidation was performed on the ores to determine whether this method of treatment was beneficial, and which parameters, if any, resulted in significantly increased gold extraction. Given below in Figure 24 are the extraction curves for baseline cyanidation tests performed on all ores. 61   Figure 24: Baseline cyanidation curves for all tested ores; from left to right: sample 165, 167, 169, 169; (35°C temperature, 20% solids by mass (initial condition), no air sparging, pH > 11) For the figures above, a brief commentary on the calculated head grades and extractions observed for each ore is given below, as well as any discussion on the trends seen in the extraction curves themselves. Table 15: Results of Baseline Cyanidation Tests on All Tested Ores Sample Measured Baseline Test #1 Baseline Test #2 Average Fire Assayed Gold Head Grade (g/mt) Calculated Gold Head Grade (g/mt) Gold Extraction (%) Calculated Gold Head Grade (g/mt) Gold Extraction (%) Gold Head Grade (g/mt) Baseline Gold Extraction (%) 165 13.5 11.05 77.8 12.7 80.9 11.8 79.5 167 19.4 18.73 86.5 17.7 86.6 18.2 85.6 169 7.57 7.16 66.8 7.70 65.0 7.28 65.9 201 4.11 4.17 81.0 4.21 81.0 4.19 81.0  0%10%20%30%40%50%60%70%80%90%100%0 5 10 15 20 25Gold Extraction %Time (Hours)165 Base. Cyan. 1165 Base. Cyan. 2Avg. Base. Cyan.0%10%20%30%40%50%60%70%80%90%100%0 5 10 15 20 25Gold Extraction %Time (Hours)167 Base. Cyan. 1167 Base. Cyan. 2Avg. Base. Cyan.0%10%20%30%40%50%60%70%80%90%100%0 5 10 15 20 25Gold Extraction %Time (Hours)169 Base. Cyan. 1169 Base. Cyan. 2Avg. Base. Cyan.0%10%20%30%40%50%60%70%80%90%100%0 5 10 15 20 25Gold Extraction %Time (Hours)201 Base. Cyan. 1201 Base. Cyan. 2Avg. Base. Cyan.62  As shown in Table 15 above, all extraction curves agree within 5% of each other, with nearly all gold lixiviation occurring within one hour of starting the cyanidation process. Gold extraction levels were calculated to be mostly complete after one or two hours of leaching, with the remaining time left for incremental increase in the amount of gold leached from the ore. This behaviour is a seen with most ores that are processed through conventional gold leaching, and the extra time needed to run these tests is deemed necessary by operators to extract the highest amount of gold from lean ores as possible. Test results for samples 165, 167, and 169 show lower calculated head grade of gold than what was initially assayed, while sample 201 produces an average calculated gold head grade that is slightly larger than the fire assay head. Where this discrepancy is larger than 1 g/mt, the difference may be due to a “nugget effect” which can allow a large grain of undigested gold to remain in the residual tailings of the ore, resulting in lower extraction values. The nugget effect could have also inflated the fire assay results of gold in the ore, for the same reason as before – a larger grain of gold may have been present in the sample sent for analysis, despite best efforts to homogenize and evenly distribute the gold throughout ore samples. For nearly every single extraction curve, a small bump in the level of gold can be seen at hour 1 for both curves, followed by a dip between hours 1 and 4, and then a gradual increase in the extraction of gold over time until the end of the tests. An explanation can be offered by means of a gradual dilution of the leachate that occurs as wash water is added to the reactor over time (at each point in the curves, e.g. at 2, 4, 8 hours). This is a result of washing the various probes used to measure physical parameters such as pH, ORP, D.O., and temperature. Gold concentrations then rise slowly, and this gradual rise may be attributed to gradual evaporation that occurred in the reactors, especially over the period from 8 hours to 24 hours. It is important to note that although the curves for extraction are fluctuating over time, the final values of extraction use the actual gold concentrations, and thus are considered the “true values” for overall gold extraction. It can be reasonably assumed then that gold extraction follows a “normal” upward trend over the first 1-2 hours of extraction, followed by a plateau where only a slight increase of gold in solution occurs. Tests involving samples 165 and 167 did not require addition of sodium hydroxide solution, either because of excess lime in the grinding stage, or because the pH conditioning stage worked as intended to neutralize any sulphuric acid that may have been generated during the grind procedure. Ores 169 and 201 did require some NaOH solution to keep pH above 11. Most ores slurries experienced a gradual decrease in pH over time as the cyanidation tests were conducted. This was most likely caused by the adsorption of carbonic acid (HCO3-), which happens as carbon dioxide from the atmosphere reacts with water in the reactor. Alternatively, some sulphur oxidation might have occurred during these tests, which produced sulphuric acid that lowered pH over time. However, with tests with prior alkaline oxidation, it is assumed that no sulphur oxidation occurs without forced aeration of the leach slurry and this theory was evidenced by 63  relatively insignificant amounts (< 10 mL) of 1.0 N NaOH solution required to keep slurries at a high pH. If significant amounts of sulphur in the ores were being oxidized even without aeration, a sharp decrease in pH would have been observed followed by rapid introduction of caustic solution. Temperatures for all trials were stable at 35.0 ± 0.5°C throughout the tests. Dissolved Oxygen (D.O.) levels were seen in almost all cases to rise over time. One possible explanation is the gradual build-up of dissolved oxygen in solution over time due to natural ingress of air at the surface of the slurry, assisted by vigorous agitation. It is important to note that all cyanidation trials avoided oxygen starvation conditions, where no dissolved oxygen is present in solution. As discussed above, gold lixiviation (shown in the Elsner equation in Chapter 2) relies on the adequate presence of dissolved or aqueous oxygen in solution, so it is important to maintain a high level of oxygen in the leach Another variable of interest for these, as well as subsequent trials, is the amount of cyanide consumed. Each trial started with approximately 10.0 kg NaCN per tonne ore and experienced a varied drop in residual cyanide, determined through volumetric titration with silver nitrate (AgNO3) previously described in Chapter 3. Given below is a summary of cyanide consumption for baseline cyanidation tests. Table 16: Cyanide Consumption for Baseline Cyanidation Tests Conducted for All Ores Sample Baseline Test #1 Baseline Test #2 Average Initial Dosage (kg/t) Cyanide Consumption (kg/t) Initial Dosage (kg/t) Cyanide Consumption (kg/t) Initial Dosage (kg/t) Cyanide Consumption (kg/t) 165 10.0 2.39 10.0 5.09 10.0 3.72 167 10.0 9.02 10.0 6.18 10.0 7.63 169 10.0 9.66 10.0 9.58 10.0 9.62 201 10.0 9.25 10.0 4.73 10.0 7.09  From the table above, sample 165, which has a significantly lower sulphur grade, exhibits lower cyanide consumption. For sample 201, which did have thiocyanate testing performed on the Pregnant Leach Solutions obtained at the end of each trial, the results are in close agreement: thiocyanate concentrations of 276 mg/L and 257 mg/L, respectively, which then translate to 1.18 and 1.17 kg/t of sodium cyanide consumed through thiocyanate production for each trial. The rest of the observed consumption can be assumed to be oxidation to cyanate (CNO-) for reasons that will be explained in detail in section 4.2.4 below. While the amount of cyanide consumed by thiocyanate production was in close agreement between the two baseline tests, there is a large discrepancy in overall cyanide consumption for sample 201 (and to a 64  lesser extent, samples 165 and 167). Several factors may influence this difference; however, the most likely scenario is that incidental dilution may have lowered the amount of residual cyanide detected through titration, thus exaggerating the calculated cyanide consumption beyond its likely true value. Recalling Figure 4 in Chapter 2 (Yannopoulos, 1991), we can see that at a high pH, there is enough electrochemical potential in the slurry to oxidize cyanide to cyanate. For sample 169, there is a major loss of cyanide and this may have significantly impacted gold extraction. No thiocyanate testing was performed for the baseline tests on sample 169, however thiocyanate levels recorded on oxidative tests show high concentrations of the compound, as will be explained below in Section 4.2.4. It appears that thiocyanate and cyanate production may severely impact the economics of gold extraction for high sulphur grade ores. It may behoove industrial operators to introduce a solid/liquid separation stage in between pre-oxidation and cyanidation stages to remove aqueous sulphur compounds such as thiosulphate from interfering with the gold extraction process. It may be that native sulphur or partially oxidized sulphur species act as a major consumer of cyanide, producing high concentrations of thiocyanate, and so it is necessary to fully oxidize these species to sulphate, which is not known to react with cyanide. However, this is merely speculation, as Ion Chromatography, or UV-Vis Spectroscopy analysis was not performed on the aqueous samples in this work. Similarly, ICP analysis on the final gold-bearing solutions from these tests would be helpful in confirming that no aqueous metal is present in solution (conclusively eliminating the possibility of metal-cyanide complexes being formed during cyanidation). Determining which thionic species are present in solution over the course of each stage may be the next research goal going forward. The consumption of cyanide via oxidation to cyanate cannot be effectively mitigated without affecting dissolved oxygen content in solution (which is necessary for the lixiviation of gold), and this loss may be a tolerated part of the gold extraction process, if the overall levels of cyanide in solution do not drop below an acceptable threshold.  4.2 Results of Alkaline Oxidation Tests  Alkaline oxidation trials were conducted using the experimental setup described in Chapter 3. These tests examined three parameters: caustic consumption, rate of air input (sparging), and oxidation temperature, to determine which of these metrics influence gold extraction the most positively. Caustic consumption was recorded as the amount of NaOH required to maintain ore slurries at a pH of 11 over the course of oxidation and cyanidation, including a pH conditioning phase if needed. Air sparging was maintained via flow meters which regulated the amount of natural air that entered the slurry at a constant pressure of 0.68 atm (10 psi). Oxidation temperature was originally measured by two external thermometers to ensure an even heat 65  distribution throughout the water bath, and the actual temperature of the slurries was recorded with a thermocouple connected to the pH/ORP meter specified in Chapter 3. Oxidation tests focused on variation of one variable at a time. Except for baseline tests (both cyanidation and oxidation), only one trial was conducted at each set of conditions, for each separate ore. Table 17 below gives a summary of all conditions used for baseline tests as well as varied parameter tests. As with Baseline Cyanidation, the methods used to calculate the levels of oxidation and gold extraction are given below in Appendix C. Example Mass Balance tables are included in Appendix D. Table 17: Summary of Oxidation and Cyanidation Conditions Used for all Ore Samples Parameter Baseline Oxidation Conditions Low High Initial Solids % 20 Oxidation Stirring Rate (RPM) 1250* Initial Caustic Dosage (kg/t) 7.5 Caustic Dosage During Oxidation (kg/t) As needed to maintain pH 11 Rate of Air Input in Litres per Minute (LPM at 0.68 atm, or 10 psi)** 2.0 1.5 2.5 Oxidation Temperature (°C) 35 35 45, 55 Cyanidation Temperature (°C) 35 Cyanidation Stirring Rate (RPM) 2000* Cyanidation Air Input (LPM) 0 Caustic Dosage During Cyanidation (kg/t) As needed to maintain pH 11  * as needed to ensure agitation at the surface of the slurry ** air pressure measured at the air tank regulator; rate of airflow measured at a flowmeter prior to air entering the air moisturizers (see Figure 23).  66  4.2.1 Caustic Consumption  For the alkaline oxidation tests conducted for this work, an initial dosage of 7.5 kg/t caustic sodium hydroxide was added to the leaching vessel at the start of oxidation. After oxidation was started, a 1.0 N (1.0 M) caustic solution was used to maintain a pH of 11. The dosage of caustic was recorded for all three stages of the test: alkaline oxidation, pH conditioning (if needed), and cyanidation. For the discussion below, the total dosage of caustic was considered for all tests, since this reference point is applicable to all tests, regardless of which stages were performed for ore samples (e.g. baseline cyanidation, without alkaline pre-oxidation, or low sulphur grade ores which did not require pH conditioning). Caustic consumption was used as a measure of how much sodium hydroxide was used to keep pH at 11 throughout oxidation and cyanidation for all trials, and is an indirect measure of both the amount of sulphuric acid that may have been evolved because of pyrite oxidation and more directly as a measure of how much hydroxide was consumed to produce oxidation products such as goethite, ferrihydrite, ferrous iron hydroxide, etc. Caustic consumption in this case would not be considered its own independent variable, but the recorded data for maintaining experimental conditions (in this case, a specific pH). For the purposes of identification, pH would be the “independent variable” that would be varied throughout a relevant alkaline range. For future testing, all the tests in this work could be repeated at different pH values to gain a better understanding of the behaviour of these iron-sulphide-containing ores. 4.2.1.1 Sample 165  Given below in Figure 25 is a plot of oxidation levels vs. total caustic dosage.  Figure 25: Extent of oxidation vs. caustic consumption for sample 165 0%1%2%3%4%5%6%7%8%9%10%7.45 7.50 7.55 7.60Extent of Oxidation (%)NaOH Dosed (kg/t)Extent of Oxidation (%)67  From the above chart, a correlation may exist between the initial dosage of caustic with oxidation. Compared to other ores however, ore sample 165 achieves an almost negligible amount of oxidation, so the above trend may merely be a product of normal experimental variation. Since sample 165 is a low sulphur ore, and 8.5 kg/t lime (Ca(OH)2) was already dosed during grinding, no further caustic was needed to maintain a high pH. The alkaline caustic solution acts to expose more of the iron sulphides in the ore to the oxidative environment inside the leaching reactor, likely by dissolving other constituents in the ore such as quartz or silicates/clays. This in turn increases the amount of iron sulphides that can be oxidized through exposure to dissolved oxygen. However, for sample 165, this is not necessarily beneficial, as will be seen from the plot of gold extraction vs. extent of oxidation below.  Figure 26: Gold extraction vs. extent of oxidation for sample 165 The above plot assumes that no oxidation occurs in tests with tests involving just a cyanidation stage (baseline cyanidation conditions). As with all results in this chapter, the extent of oxidation and extraction of gold was calculated using the methods described in Appendix C. The two data for gold extractions under baseline cyanidation conditions are therefore plotted on the 0% oxidation line. A maximum gold extraction is seen at approximately 4% sulphur oxidation. The two highest points in the plot above correspond to the gold extractions observed from baseline oxidation conditions (35°C oxidation temperature, 2.0 LPM air sparging at 0.68 atm pressure), and gold extractions steadily decrease from this peak. A possible explanation could include partial sulphide oxidation to native sulphur, which could coat ore particles and prevent cyanide leaching solution from contacting gold grains. Alternatively, iron oxidation, dissolution, and immediate reprecipitation in alkaline solution could form amorphous iron oxide or -hydroxide surface layers, which could also act to prevent cyanide leaching. Either the morphology of the 0%10%20%30%40%50%60%70%80%90%100%0% 2% 4% 6% 8%Gold Extraction (%)Extent of Oxidation (%)Gold Extraction (%)68  coating or the thickness of the surface layer could prevent proper ingress of cyanide leachate, and this could hinder gold extractions as extent of oxidation increases. It is also possible that some aqueous ferrous iron remained in solution to form iron-cyanide complexes; however, this is not very likely as aqueous iron levels detected in samples from the oxidation phase showed extremely low concentrations of iron (<5 mg/L). In future, ICP testing could be performed on aqueous samples from the gold leaching stage to confirm this explanation. Again, examining gold extraction versus caustic consumption yields the same trends, as will be shown in Figure 27 and Figure 28 below:  Figure 27: Gold extraction vs. caustic consumption for sample 165; alkaline oxidation tests results only  Figure 28: Gold extraction vs. caustic consumption for sample 165; baseline cyanidation results included 0%10%20%30%40%50%60%70%80%90%100%7.45 7.50 7.55 7.60Gold Extraction (%)NaOH Dosed (kg/t)Gold Extraction (%)0%10%20%30%40%50%60%70%80%90%100%0 2 4 6 8Gold Extraction (%)NaOH Dosed (kg/t)Gold Extraction (%)69  When considering just the alkaline oxidation tests, caustic consumption does not appear to increase the level of gold extraction; when baseline cyanidation tests are included, however, gold extraction increases at higher caustic dosages compared to no initial dosage of caustic. The increase in sodium hydroxide levels may be exposing more gold to cyanide solution by dissolution of silicates, clays, and other gangue minerals in the ore. The increase in gold extraction could also be attributed to increased iron and sulphur oxidation due to air sparging, which increases the dissolved oxygen content in solution and thus promotes oxidation of sulphide ores. Further discussion on the level of air sparging will be given below in Section 4.2.2. 4.2.1.2 Sample 167  Given below is a plot of the extent of oxidation vs. caustic consumption for Sample 167:  Figure 29: Extent of oxidation vs. caustic consumption for sample 167 A very linear trend emerges in the above plot. This could indicate a causal relationship between the amount of caustic dosed over the course of each test and the final extent of oxidation that is achieved. Since caustic sodium hydroxide was automatically dosed whenever pH dropped below 11, this graph shows that more sulphuric acid is generated as a greater extent of oxidation is achieved. Additionally, it may also indicate that caustic hydroxide (OH-) is consumed as oxidation by-products such as goethite (FeOOH) is formed. The lowest point on the graph resulted from baseline oxidation conditions, whereas the two highest points were the result of high and medium temperature (55°C and 45°C, respectively). In contrast to sample 165, which showed a somewhat adverse effect of oxidation on gold extraction, sample 167 shows a marked improvement after pre-oxidative treatment. Given below in Figure 30 is a plot of gold extraction vs. extent of oxidation for sample 167. 0%10%20%30%40%50%60%70%80%90%100%25 30 35 40 45Extent of Oxidation (%)NaOH Dosed (kg/t)Extent of Oxidation (%)70   Figure 30: Gold extraction vs. extent of oxidation for sample 167; baseline cyanidation data lie at 0% oxidation For the graphs in this section, all tests conducted for this research are considered, despite varying conditions for rates of airflow and oxidation temperatures. This is because while those parameters are considered independent variables, caustic consumption is not such a variable. Here, caustic consumption indicates an increase in the amount of sulphuric acid evolved from pyrite/marcasite oxidation, as well as production of oxidation products which may consume the hydroxide (OH-) ion. As previously mentioned, the increased presence of caustic in solution may expose more iron sulphides to an oxidative environment by attacking or dissociating other constituents such as quartz or other silicates and clays in the ore. As will be discussed below in Section 4.4, caustic consumption is a measure of the amount of sodium hydroxide dosed to achieve a specific pH. In future, more testing may be carried out with other pH values, to expand upon the knowledge presented in this work. Baseline cyanidation is included and is assumed to have very little, if no prior oxidation and so therefore lies on the line x = 0. This is evidenced by the extremely low consumption of caustic sodium hydroxide (comparatively speaking) during baseline cyanidation (tests conducted without pre-aeration). For those tests, pH control did not have to be implemented nearly as often, and so lower amounts of caustic was not dosed during those tests, since less sulphuric acid was released from the ore and a smaller amount of iron hydroxide oxidation products were formed (consuming fewer hydroxide ions). The highest gold extractions are again associated with high temperature tests. Finally, a clear correlation can be seen in the graph of gold extraction vs. caustic sodium hydroxide dosed, given in Figure 31 below. 0%10%20%30%40%50%60%70%80%90%100%0% 20% 40% 60%Gold Extraction (%)Extent of Oxidation (%)Gold Extraction (%)71   Figure 31: Gold extraction vs. caustic consumption for sample 167; baseline cyanidation results included Increased oxidation of sulphide ore due to several factors (including forced aeration of the slurry, elevated temperatures, and caustic sodium hydroxide dissolving mineral particles) exposes more gold to the cyanide solution in the reactor. The increased levels of caustic are a by-product of the choice to maintain a high pH in the reactors. In addition, the caustic in the slurry may be dissolving gangue constituents of the ore, thereby aiding in the exposure of more iron sulphides to an oxidative environment and further exposing more gold to cyanide solution during gold leaching. It may be possible to determine to what extent caustic solution is dissolving ore without iron sulphide oxidation if tests are performed under anaerobic conditions (e.g. under a nitrogen or argon atmosphere) without air sparging, as it is expected that no oxidation would occur under such conditions without some oxidant in solution. 4.2.1.3 Sample 169  The plot of extent of oxidation vs. caustic dosage for sample 169 is given below: 0%10%20%30%40%50%60%70%80%90%100%15 25 35 45Gold Extraction (%)NaOH Dosed (kg/t)Gold Extraction (%)72   Figure 32: Extent of oxidation vs. caustic consumption for sample 169 A close grouping of data is present on the left side of the graph above, and a single data point is present on the right. This point belongs to the high temperature (Test F), which required approximately 40.4 kg/t NaOH to maintain a pH of 11 through all stages. The second highest point on the graph is the medium temperature test (Test E), which yielded over 50% oxidation, but did not require as much caustic. This may be due to the relatively low amounts of goethite and ferrihydrite observed for this ore; less caustic may be consumed with production of oxidation products such as magnetite and hematite, compared to goethite and ferrous or ferric hydroxides. Given below are the plots of gold extraction vs. extent of oxidation and gold extraction vs. caustic consumption, in Figure 33 and Figure 34, respectively.  Figure 33: Gold extraction vs. extent of oxidation for sample 169; baseline cyanidation results included 0%10%20%30%40%50%60%70%80%90%100%25 30 35 40 45Extent of Oxidation (%)NaOH Dosed (kg/t)Extent of Oxidation (%)0%10%20%30%40%50%60%70%80%90%100%0% 20% 40% 60%Gold Extraction (%)Extent of Oxidation (%)Gold Extraction (%)73  For the plot above, a moderate increase is seen with respect to gold extraction with and without oxidation. Again, it assumed that no sulphur oxidation occurs during tests without aeration (air sparging). These tests yielded an average of 66% of the gold in the ore samples. With oxidation, gold extractions increased to above 71% to just over 73%, a small margin. As with ore 167, this sample had a high sulphur grade and responded positively to oxidizing some sulphur so that more gold could be exposed to cyanide solution.  Figure 34: Gold extraction vs. caustic consumption for sample 169; baseline cyanidation results included Like Figure 33 above, the plot of gold extraction vs. caustic consumption for sample 169 shows an increase in gold extraction after pre-oxidation, and specifically with an increased presence of caustic. 4.2.1.4 Sample 201  Given below in Figure 35 is the plot of extent of oxidation vs. caustic consumption. 0%10%20%30%40%50%60%70%80%90%100%15 20 25 30 35 40 45Gold Extraction (%)NaOH Dosed (kg/t)Gold Extraction (%)74   Figure 35: Extent of oxidation vs. caustic consumption for sample 201 Sample 201 shows an improvement in the oxidation of the ore with an increase of caustic consumption. The fit for all data is quite good, and the highest data were the results from high and medium temperature tests, and the lowest data were the results of either baseline oxidation or low air input tests. The plot of gold extraction vs. extent of oxidation is given below:  Figure 36: Gold extraction vs. extent of oxidation for sample 201; baseline cyanidation assumed to have no oxidation Sample 201 shows a slight improvement in gold extraction with increasing extent of oxidation. The lowest data point belonged to the low-air oxidation test, which may have been affected by a large assayed value of gold in the tails residue, despite aqueous gold concentrations that were close to other tests. Again, the 0%10%20%30%40%50%60%70%80%90%100%15 20 25 30 35Extent of Oxidation (%)NaOH Dosed (kg/t)Extent of Oxidation (%)0%10%20%30%40%50%60%70%80%90%100%0% 10% 20% 30% 40% 50%Gold Extraction (%)Extent of Oxidation (%)Gold Extraction (%)75  highest extraction seen here was the result of high temperature oxidation. Finally, the plot of gold extraction vs. caustic consumption for sample 201 is given below in Figure 37:  Figure 37: Gold extraction vs. caustic consumption for sample 201; baseline cyanidation results included For this ore, it was difficult to fit the final extractions according to a linear trendline due to the scatter of gold extraction data, despite nearly all tests showing favourable results. Generally, gold extraction is seen to increase as higher dosages of caustic are introduced, again most likely due to increased exposure of iron sulphide minerals to the oxidative environment because of dissolution of other minerals in the ore. 4.2.2 Air Sparging  The analyses for gold extraction in relation to air sparging are the most difficult to rationalize, due to the experimental error associated with maintaining constant airflow to the leaching vessels during oxidation. It was difficult to accurately control air sparging rates (and oxygen uptake into solution) because the glass frits used to pump air into the slurries kept getting clogged by small ore particles (<50 μm). This would eventually lead to a build up of pressure in the air moisturizers (see Figure 23) which would eventually release that pressure by popping their rubber stopper lids. After this point, it would be almost impossible to maintain a constant desired level of air input to the reactors. Only by repeatedly taping the lids of the Erlenmeyer flasks down, a desired level of air flow achieved. Even after this measure, sporadic drops in pressure were observed throughout some oxidation tests. Glass frits were changed after every test and cleaned with a concentrated acid solution after each test before being thoroughly rinsed and reused. Despite this, problems with air flow were sometimes noted. The use of a tube to aerate slurries would be considered 0%10%20%30%40%50%60%70%80%90%100%10 15 20 25 30 35 40Gold Extraction (%)NaOH Dosed (kg/t)Gold Extraction (%)76  for future tests, if the problem of large air bubbles could be adequately solved – a tube may not be able to aerate the solutions with small enough air bubbles while still maintaining a high rate of airflow. Generally, it is expected that an increased flow of air into an ore slurry would result in an increase in dissolved oxygen. Higher D.O. levels are expected to increase the extent of oxidation of sulphur and iron, since oxygen is thought to be the main oxidant of iron sulphide minerals in this scenario. However, too much airflow into the slurry can allow large bubbles to form in the leachate and this could lead to spattering of ore solids against the reactor walls, above the level of the ore slurry. While this spattering may not seem a significant contributor to lowered oxidation levels and gold extractions, every effort was taken to ensure a uniform residence time of all ore solids in the leaching solution. A wash of the reactor walls would be performed using DI water during one of the sampling times. In addition to the problems previously described, large bubbles formed by excessive airflow to the reactors could potentially result in lower dissolved oxygen values, since uptake of oxygen into solution is dependent on surface area contact with air bubbles. Some discussion related to each ore specifically is given below. 4.2.2.1 Sample 165  Given below in Figure 38 is a plot of the extent of oxidation achieved for sample 165 against rate of air input. All data points are close to each other and can be considered the same within normal experimental variation (i.e. the percent difference between them is below 2%). Oxidation rises overall as more air sparging is maintained in the leaching reactor. As previously stated, for ore 165 an increasing extent of oxidation may hinder gold extraction, due to possible precipitation of iron oxide surface layers.  Figure 38: Extent of oxidation vs. rate of air input at 0.68 atm (10 psi) for sample 165 0%1%2%3%4%5%6%7%8%9%10%1 1.5 2 2.5 3Extent of Oxidation (%)Rate of Air Input at 0.68 atm (10 psi) Pressure (LPM)Extent of Oxidation (%)77  As shown in below in Figure 39, forced air sparging into the leaching reactor does marginally help to increase gold extraction. This is illustrated below in Figure 39. A slight decrease is seen in the extraction of gold with air sparging rates over 2.0 LPM, which may indicate either poor oxygen uptake due to large air bubbles being formed, or unfavourable precipitation of oxidation products, as discussed above.  Figure 39: Gold extraction vs. rate of air input at 0.68 atm (10 psi) pressure for sample 165; baseline cyanidation results included 4.2.2.2 Sample 167  Given below in Figure 40 is a plot of extent of oxidation vs. rate of air input at constant 0.68 atm (10 psi) pressure for sample 167. The trend is like that observed for sample 165: 0%10%20%30%40%50%60%70%80%90%100%0 0.5 1 1.5 2 2.5 3Gold Extraction (%)Rate of Air Input at 0.68 atm (10 psi) Pressure (LPM)Gold Extraction (%)78   Figure 40: Extent of oxidation vs. rate of air input at 0.68 atm (10 psi) for sample 167 An overall increase in oxidation is observed when increasing the rate of air input from 2.0 LPM to 2.5 LPM, with a slight dip at the 2.0 LPM mark. Regardless of the shape of the curve, the same trend may result in an increase of gold extraction compared with baseline cyanidation, as shown in Figure 41 below:  Figure 41: Gold extraction vs. rate of air input at 0.68 atm (10 psi) pressure for sample 167; baseline cyanidation results included Overall a very slight increase in gold extractions is seen when compared with baseline cyanidation, which had no forced air sparging. Baseline cyanidation gold extraction levels for ore 167 were approximately 85.6% and increased to just over 90% with air sparging in slurry during the oxidation stage. While this may 0%10%20%30%40%50%60%70%80%90%100%1 1.5 2 2.5 3Extent of Oxidation (%)Rate of Air Input at 0.68 atm (10 psi) Pressure (LPM)Extent of Oxidation (%)0%10%20%30%40%50%60%70%80%90%100%0 0.5 1 1.5 2 2.5 3Gold Extraction (%)Rate of Air Input at 0.68 atm (10 psi) Pressure (LPM)Gold Extraction (%)79  seem like only a small difference, with the high value of gold metal, this result could potentially be worth a lot of increased value extracted from this type of ore. A plateau and even a slight decrease is seen after 2.0 LPM air input, but this appears to be within normal experimental variation. Given below is a plot of extents of oxidation vs. dissolved oxygen in solution during oxidation.  Figure 42: Extent of oxidation vs. average dissolved oxygen in solution during oxidation for sample 167 A clear decrease in extent of oxidation with increasing average dissolved oxygen in solution is seen in Figure 42 above. This trend can be explained by the fact that increased oxidation of iron sulphide minerals consumed more oxygen in solution, while lower extents of oxidation were achieved with a proportionally lower consumption of D.O. While air was being sparged into the system, the low oxygen uptake into solution and high oxidation rates meant that dissolved oxygen could not be replenished fast enough to maintain a constant level of D.O. The average concentration recorded of oxygen in solution throughout pre-treatment was thus decreased as a result. 4.2.2.3 Sample 169  For sample 169, rates of air input have a moderate effect on extents of oxidation, as well as on gold extraction, as will be shown in the two graphs below: 0%10%20%30%40%50%60%70%80%90%100%0 1 2 3 4 5Extent of Oxidation (%)Dissolved Oxygen in Solution (mg/L)Extent of Oxidation (%)80   Figure 43: Extent of oxidation vs. rate of air input at 0.68 atm (10 psi) for sample 169 A decrease is observed as air sparging rates increase from 1.5 LPM to 2.0 LPM, from over 40% to just above 35% oxidation, and a plateau is seen as airflow increases further to 2.5 LPM. Generally, increased dissolved oxygen levels in solution should increase the amount of sulphur and iron oxidized in the ore. However, increasing rates of airflow may not actually increase dissolved oxygen levels to a significant degree – due to either large bubbles being formed in solution, which hinder oxygen uptake into the solution, or lead to spattering of ore solids onto the reactor walls above the level of leaching solution. While the temporary loss of solids may not be a significant issue, this leads to an overall reduced average residence time in the leaching solution. The problem is exacerbated by the sticky, viscous nature of the ores in caustic solution; significant amounts of ore solids can accrue above the level of leaching solution without periodic washing of the reactor walls (e.g. leaching overnight). Finally, as mentioned before, it was difficult to maintain air flow and pressure to each leaching vessel, especially at higher air input rates, because tiny ore particles tended to clog the glass frits that were used to bubble air into the slurry. 0%10%20%30%40%50%60%70%80%90%100%1 1.5 2 2.5 3Extent of Oxidation (%)Rate of Air Input at 0.68 atm (10 psi) Pressure (LPM)Extent of Oxidation (%)81   Figure 44: Gold extraction vs. rate of air input at 0.68 atm (10 psi) for sample 169; baseline cyanidation results included A large increase in gold extraction is observed with increased air sparging rates, especially for ores with higher sulphur grades such as samples 167 and 169. Overall, air sparging during a pre-treatment phase is seen to be an effective way of increasing gold extraction for this ore. Given below in Figure 45 is a plot of extent of oxidation vs. dissolved oxygen in solution during oxidation. The same downward trend seen with ore sample 167 (discussed above) and explanation is offered here.   Figure 45: Extent of oxidation vs. average dissolved oxygen in solution during oxidation for sample 169 0%10%20%30%40%50%60%70%80%90%100%0 1 2 3Gold Extraction (%)Rate of Air Input at 0.68 atm (10 psi) Pressure (LPM)Gold Extraction (%)0%10%20%30%40%50%60%70%80%90%100%0 1 2 3 4Extent of Oxidation (%)Dissolved Oxygen in Solution (mg/L)Extent of Oxidation (%)82  4.2.2.4 Sample 201  Given below is a plot of extent of oxidation vs. rate of air input for sample 201:  Figure 46: Extent of oxidation vs. rate of air input at 0.68 atm (10 psi) pressure for sample 201 Sample 201 shows a slight upward trend with regards to sulphur oxidation compared to rates of air sparging, shown in the graph above. As the rate of air flow into the leach reactor increases from 1.5 LPM to 2.0 LPM, oxidation sharply increases, while oxidation plateaus (within normal experimental variation) after that. The total difference between the highest and lowest oxidation observed in the above graph is less than 2%, so it may be that this trend is not significant at all. More trials are needed to accurately determine whether this trend is statistically reliable. An increase of this small magnitude may not significantly impact gold extraction, since the interval between tested air rates was only 0.5 LPM – possibly not enough to reveal a meaningful behaviour in the ores tested for this work, given the experimental error for these trials. Given below is the plot of gold extraction vs. rate of air input, with baseline cyanidation (without pre-treatment) plotted on the x = 0 line. 0%10%20%30%40%50%60%70%80%90%100%1 1.5 2 2.5 3Extent of Oxidation (%)Rate of Air Input at 0.68 atm (10 psi) Pressure (LPM)Extent of Oxidation (%)83   Figure 47: Gold extraction vs. rate of air input at 0.68 atm (10 psi) pressure for sample 201; baseline cyanidation results included Gold extraction does increase with the addition of extra air sparging during oxidation. This seems to confirm the trend seen in Figure 46, which shows that oxidation increases with the addition of more air into the leach slurry. The total increase from baseline cyanidation to the maximum gold extraction (resulting from high temperature oxidation) is approximately 10%, which is a substantial amount. A plot relating extent of oxidation vs. average dissolved oxygen in solution during oxidative pre-treatment is given below.  Figure 48: Extent of oxidation vs. average dissolved oxygen in solution during oxidation for sample 201 0%10%20%30%40%50%60%70%80%90%100%0 1 2 3Gold Extraction (%)Rate of Air Input at 0.68 atm (10 psi) Pressure (LPM)Gold Extraction (%)0%10%20%30%40%50%60%70%80%90%100%0 1 2 3 4 5Extent of Oxidation (%)Dissolved Oxygen in Solution (mg/L)Extent of Oxidation (%)84  A decreasing trend with respect to D.O. in solution is again observed. Since both sulphur oxidation and gold lixiviation require dissolved oxygen, it may be that a lower oxygen value in solution indicates that these reactions are occurring rapidly, consuming dissolved oxygen faster than it can be replaced. Alternatively, these trends could be indicative of low dissolved oxygen in solution at high temperatures, which promote oxidation despite the lower levels of oxygen in solution. 4.2.3 Oxidation Temperature  Oxidation temperature seemed to influence extents of oxidation and gold extractions the most strongly out of all variables considered. Refractory sulphide ores generally achieved high levels of oxidation when exposed to higher temperatures and the rates of sulphur release seen from these tests were much faster, despite sulphur oxidation being an exothermic process. The temperatures of slurries were generally very closely controlled over each test, often within 0.5°C. Any variation or scatter in the data could thus be attributed to other factors, most likely variations in rates of air sparging (or air pressure), and dilution of the slurry over time, as has previously been discussed. Cyanidation stages of all tests were held at a constant 35°C temperature, and so would not influence any individual tests. Since most trials did not require much caustic addition after the oxidation and pH conditioning phases, it can be inferred that no further significant sulphide oxidation was taking place during the gold leaching cyanidation phase. A brief discussion of the behaviour of each ore with respect to oxidation temperature is given below. 4.2.3.1 Sample 165  A plot of extent of oxidation vs. oxidation temperature is given below in Figure 49. 85   Figure 49: Extent of oxidation vs. oxidation temperature for sample 165 All alkaline oxidation tests were considered when creating the above chart, even the tests with lower or higher rates of air input (in this chart, the two highest oxidation points at 35°C). A downward trend can be seen, and this trend is exaggerated if only the baseline oxidation points are considered for the 35°C oxidation temperature (the two lowest points at 35°C) rather than those oxidation tests conducted with varying airflows. However, considering that all the results for the trials conducted with ore 165 are within a few percentage points of each other, it may be that no clear trend can be discerned with normal experimental variation. In this ore, sulphur grade is quite low (0.88%, assayed), and so this may exaggerate the severity of variable interaction with oxidation and gold extraction results, since a wild swing in these metrics can be achieved with a rather small or insignificant change in aqueous sulphur or gold in solution (or perhaps a nugget of unexposed material that is left in the tails residues). A plot of gold extraction vs. oxidation temperature is given below in Figure 50. 0%1%2%3%4%5%6%7%8%9%10%30 35 40 45 50 55 60Extent of Oxidation (%)Temperature (°C)Extent of Oxidation (%)86   Figure 50: Gold extraction vs. oxidation temperature for sample 165 In the above graph, the two lowest points for the 35°C temperature are the gold extractions for the varied air input tests. The highest levels of gold extraction are seen in baseline oxidation, at 35°C, with an average gold extraction of 83% at that temperature. Gold extraction decreases slightly to approximately 81.7% at 45°C, and finally increases slightly to 82.5% at 55°C. Overall, gold extraction from ore 165 is not significantly influenced by the temperature held in the oxidation phase. This result may be reflective of the decreased oxidation levels seen for sulphide oxidation, so less gold may be exposed to cyanide solution in the gold leaching stage. A slight decrease in final gold extractions is seen after tests with elevated temperatures, but this decrease does not appear to be significant. More testing may be required to adequately explain the contrarian nature of sample 165, which did not follow the trends seen with any of the other ores tested in this work. 4.2.3.2 Sample 167  A plot of extent of oxidation vs. oxidation temperature is given below in Figure 51: 0%10%20%30%40%50%60%70%80%90%100%30 40 50 60Gold Extraction (%)Oxidation Temperature (°C)Gold Extraction (%)87   Figure 51: Extent of oxidation vs. oxidation temperature for sample 167 The above plot includes tests with varying air input at the 35°C mark (Tests C, and D, or low and high air flow, respectively). The closeness of fit of these data is still relatively high. A dramatic upward trend in oxidation is observed, and the gold extraction achieved for Sample 167 responds positively with enhanced oxidation, as will be seen below in Figure 52. Gold extraction increases from 89.5% to just over 91.8%, a slight difference. It seems that most of the tested refractory sulphide ores respond positively to increased temperature, both in terms of oxidation as well as gold extractions.  Figure 52: Gold extraction vs. oxidation temperature for sample 167 0%10%20%30%40%50%60%70%80%90%100%30 35 40 45 50 55 60Extent of Oxidation (%)Temperature (°C)Extent of Oxidation (%)0%10%20%30%40%50%60%70%80%90%100%30 40 50 60Gold Extraction (%)Oxidation Temperature (°C)Gold Extraction (%)88  4.2.3.3 Sample 169  Sample 169 responds extremely well to an increase in oxidation temperature, as shown below in Figure 53.  Figure 53: Extent of oxidation vs. oxidation temperature for sample 169 An increase of approximately 20% is seen in the extent of oxidation with an increase of 20°C in oxidation temperature. The sulphides that comprise sample 169 are extremely reactive, and despite the exothermic nature of sulphur oxidation, the increased temperature of the leach slurry results in higher extent of oxidation over the same period. However, this increase in oxidation does not appear to improve gold extraction with sample 169, as shown below. 0%10%20%30%40%50%60%70%80%90%100%30 35 40 45 50 55 60Extent of Oxidation (%)Oxidation Temperature (°C)Extent of Oxidation (%)89   Figure 54: Gold extraction vs. oxidation temperature for sample 169 Overall, the trend for gold extraction remains steady within normal experimental variation with oxidation temperature in the pre-treatment phase. A maximum of 73.1% gold extraction was achieved at the highest temperature tested (55°C) compared to a minimum of 71.8% at 45°C, and just under 72.4% at baseline temperature. It appears that despite a major increase in sulphur oxidation with increased temperature, the extraction of gold does not change significantly. Even at the highest oxidation temperature, approximately 73% of the gold in the ore is extracted. This is still a low extraction for a gold ore, indicating that despite extensive sulphur oxidation, another issue may be hampering gold extraction. Sample 169 may be a “double refractory” ore, which may have other issues, such as silicate occlusion of gold grains, or perhaps the sulphur or iron that is oxidized in the ore remains only partially dissolved, or even re-precipitates on the surface of gold-bearing mineral particles as surface product layers, thus interfering with contact with cyanide leaching solution. More tests with this type of ore is needed, with perhaps some optical, or Scanning Electron Microscopy to determine what is preventing higher gold extraction levels. 4.2.3.4 Sample 201  Given below in Figure 55 and Figure 56 are plots of extent of oxidation vs. oxidation temperature and gold extraction vs. oxidation temperature, respectively. 0%10%20%30%40%50%60%70%80%90%100%30 40 50 60Gold Extraction (%)Oxidation Temperature (°C)Gold Extraction (%)90   Figure 55: Extent of oxidation vs. oxidation temperature for sample 201  Figure 56: Gold extraction vs. oxidation temperature for sample 201 Sample 201 shows marked improvements in both oxidation as well as gold extraction as oxidation temperature increases. High and medium temperature tests yielded 90.7% and 83.0% gold extraction, respectively. Tests conducted at baseline oxidation temperatures (35°C) also yielded results in this range. However, a dramatic decrease is seen with the gold extraction of the low-air test, which may be due to the gold particle sparsity (“nugget”) effect, which may have decreased gold extraction at least 10% compared to other tests performed at the same temperatures. 0%10%20%30%40%50%60%70%80%90%100%30 35 40 45 50 55 60Extent of Oxidation (%)Temperature (°C)Extent of Oxidation (%)0%10%20%30%40%50%60%70%80%90%100%30 40 50 60Gold Extraction (%)Oxidation Temperature (°C)Gold Extraction (%)91  It is not totally clear why higher oxidation temperatures lead to increased gold extractions, despite sulphur oxidation being an exothermic process. A likely explanation is that elevated temperatures improve the initial kinetics of oxidation reactions, and this rapid sulphide oxidation exposes more gold to cyanide leaching solution earlier in the cyanidation trial, which increases the total amount of gold extracted from this ore. At higher temperatures a much more porous iron oxide or hydroxide layer may be able to precipitate on the surface of the ore compared to surface by-product layers that occur at lower temperatures. These surface layers with higher porosities allow cyanide to more easily penetrate mineral particles and leach gold.  4.2.4 Cyanide Consumption  The last parameter discussed in this chapter is cyanide consumption. Volumetric cyanide titration was carried out for all pregnant leach solution (PLS) samples, to determine the final concentration of unreacted sodium cyanide present in the leaching reactor at the end of the cyanide leaching stage. The difference between the calculated initial and final concentrations of sodium cyanide is considered as cyanide loss. Volumetric titration of cyanide is a well-established process that is carried out in industry to ensure an adequate level of cyanide during gold leaching. However, it is possible for this procedure to overestimate the amount of residual cyanide in solution. Weak Acid Dissociable (WAD) cyanide complexes (e.g. metal cyanide complexes, discussed below) can dissociate and reintroduce cyanide into the PLS, inflating the amount of cyanide left for gold lixiviation. To confirm that no WAD cyanides are present, ICP testing may be utilized for future work, and additionally any aqueous metal in solution can be quantified. Typically, consumption stems from four sources: - lixiviation with precious metals (gold and silver, typically), - formation of other metal-cyanide complexes such as iron- or copper cyanides, - formation of cyanates (CNO-) - formation of thiocyanates (SCN-) The consumption of cyanide due to precious metal lixiviation is extremely low when dealing with natural ores. The loss associated with gold- or silver-cyanide complexes is usually less than 0.05% of the total cyanide loss. Metal-cyanide complexes can potentially consume a lot of cyanide in leaching solutions in some cases. However, because oxidation was carried out at high pH, the amount of aqueous iron or copper in solution during oxidation is quite low. Iron is seen to precipitate quickly as iron oxides and -hydroxides such as hematite or goethite at high pH. This hypothesis can be developed by looking at the leaching conditions used in experiments and the iron species found in the same conditions in Pourbaix diagrams 92  from Figure 11 to Error! Reference source not found.. The presence of such iron oxide and hydroxide species is confirmed by XRD analysis of the tails residues, discussed in Section 4.3 below. Some iron may dissolve during cyanidation to form ferrous cyanide complexes. ICP testing may be carried out in future to examine the extent of cyanide consumption due to these compounds. Additionally, the results of elemental analysis for metals in each ore show that most other metal elements are in the ppm range, which is quite insignificant, so the consumption of sodium cyanide by production of other metal cyanide species is unlikely to be significant. This leaves the last two methods of cyanide consumption: oxidation to cyanates and formation of thiocyanates. For ores 169 and 201, all alkaline oxidation tests had thiocyanate testing (using Spectrophotometric Colourimetry). Sample 201 had thiocyanate testing performed on the PLS for baseline cyanidation tests as well. Sample 167 only had the PLS from its air-variation tests (Tests C and D) analyzed for thiocyanates. For sample 165, no thiocyanate testing was performed, but this might be superfluous, since the sulphur content of sample 165 is quite low, so it can be assumed that no significant thiocyanate formation occurs during cyanidation. Discussion of each ore’s cyanide consumption is given below. 4.2.4.1 Sample 165  A summary of the total consumption of cyanide for sample 165 is given below.  Table 18: Total Cyanide Consumption and Breakdown for Sample 165  NaCN Total Consumption NaCN Consumption Breakdown Test NaCN dose (kg/t) Final NaCN level (kg/t) NaCN consumed (kg/t) NaCN (Au) (kg/t) NaCN (SCN) (kg/t) NaCN (other) (kg/t) Base. Cyan. #1 10.0 7.62 2.39 2.14E-03 - - Base Cyan. #2 10.0 4.91 5.09 2.55E-03 - - Avg. Base. Cyan. 10.0 6.29 3.72 2.34E-03 - - Base. Ox. #1 (Test A) 10.0 6.15 3.89 3.30E-03 - - Base. Ox. #2 (Test B) 10.0 6.23 3.81 3.25E-03 - - Avg. Base. Ox. 10.0 6.19 3.85 3.28E-03 - - Low Air (Test C) 9.14 4.77 4.37 2.57E-03 - - High Air (Test D) 10.2 2.65 7.53 2.54E-03 - - Medium Temp. (Test E) 9.94 4.35 5.59 2.62E-03 - - 93  High Temp. (Test F) 10.4 2.65 7.74 2.69E-03 - -  As shown in all the tables in this section, NaCN consumption related to lixiviation with gold is negligible. Less than 0.01 kg/t of cyanide is consumed via extraction of any precious metals. For sample 165, the highest levels of NaCN consumption is seen after high airflow and high temperature tests. As seen above in the difference between average baseline oxidation and average baseline cyanidation (without pre-treatment), the addition of aeration during the oxidation phase slightly increases cyanide consumption.  4.2.4.2 Sample 167  The summary for cyanide consumption for ore 167 is given below: Table 19: Total Cyanide Consumption and Breakdown for Sample 167  NaCN Total Consumption NaCN Consumption Breakdown Test NaCN dose (kg/t) Final NaCN level (kg/t) NaCN consumed (kg/t) NaCN (Au) (kg/t) NaCN (SCN) (kg/t) NaCN (other) (kg/t) Base. Cyan. #1 10.0 0.98 9.02 4.03E-03 - - Base Cyan. #2 10.0 3.83 6.18 3.81E-03 - - Avg. Base. Cyan. 10.0 2.38 7.63 3.93E-03 - - Base. Ox. #1 (Test A) 10.1 1.89 8.15 3.97E-03 - - Base. Ox. #2 (Test B) 9.99 3.34 6.65 3.91E-03 - - Avg. Base. Ox. 10.0 2.63 7.39 3.94E-03 - - Low Air (Test C) 10.1 2.17 7.94 3.91E-03 3.47 4.47 High Air (Test D) 10.1 3.96 6.15 3.87E-03 4.41 1.73 Medium Temp. (Test E) 10.2 2.51 7.70 4.10E-03 - - High Temp. (Test F) 10.1 2.38 7.73 4.22E-03 - -  As shown in the table above, sample 167 has a much higher overall level of cyanide consumption. There is not much difference between baseline cyanidation and tests with oxidative pre-treatment before gold leaching. For the two tests which received thiocyanate testing on their PLS samples, thiocyanate levels were increased slightly with a higher rate of air flow during oxidation.  94  4.2.4.3 Sample 169  The summary for cyanide consumption for ore 169 is given below: Table 20: Total Cyanide Consumption and Breakdown for Sample 169  NaCN Total Consumption NaCN Consumption Breakdown Test NaCN dose (kg/t) Final NaCN level (kg/t) NaCN consumed (kg/t) NaCN (Au) (kg/t) NaCN (SCN) (kg/t) NaCN (other) (kg/t) Base. Cyan. #1 10.0 0.35 9.66 1.19E-03 - - Base Cyan. #2 10.0 0.43 9.58 1.20E-03 - - Avg. Base. Cyan. 10.0 0.39 9.62 1.19E-03 - - Base. Ox. #1 (Test A) 9.98 1.69 8.29 1.31E-03 3.85 4.44 Base. Ox. #2 (Test B) 9.89 3.72 6.17 1.28E-03 5.28 0.90 Avg. Base. Ox. 9.93 2.66 7.27 1.29E-03 4.55 2.72 Low Air (Test C) 10.0 1.34 8.67 1.34E-03 3.70 4.97 High Air (Test D) 9.91 4.84 5.07 1.31E-03 0.99 4.08 Medium Temp. (Test E) 10.0 1.07 8.92 1.32E-03 2.60 6.32 High Temp. (Test F) 10.1 0.90 9.20 1.35E-03 2.66 6.54  Sample 169, like the previous sample, shows high levels of cyanide consumption after gold leaching. Although cyanide consumption is high, this cannot adequately explain the low gold extraction levels seen for this ore – as other ores also experience similarly high levels of cyanide consumption, but still realise higher levels of extraction through cyanidation. Thiocyanate testing was performed on almost all PLS samples for this ore, with the highest levels of SCN detected for tests with baseline oxidation. SCN levels were much lower for medium and high temperature tests, indicating that while sulphur oxidation was high for these tests, thiocyanate levels might have been negatively affected by relatively lower amounts of partially oxidized sulphur in solution. At higher temperatures, sulphur may have been fully oxidized to sulphate, which is non-reactive. Cyanide consumption was high for most tests, but none of the tests had negligible cyanide in the final solutions. It is unlikely that cyanide destruction was the cause of the lower gold extractions seen for this ore.  95  4.2.4.4 Sample 201  Finally, the cyanide consumption for ore 201 is given below: Table 21: Total Cyanide Consumption and Breakdown for Sample 201  NaCN Total Consumption NaCN Consumption Breakdown Test NaCN dose (kg/t) Final NaCN level (kg/t) NaCN consumed (kg/t) NaCN (Au) (kg/t) NaCN (SCN) (kg/t) NaCN (other) (kg/t) Base. Cyan. #1 10.0 0.75 9.25 8.41E-04 1.18 8.07 Base Cyan. #2 10.0 5.27 4.73 8.49E-04 1.17 3.56 Avg. Base. Cyan. 10.0 2.92 7.09 8.46E-04 1.17 5.92 Base. Ox. #1 (Test A) 10.2 1.68 8.48 9.32E-04 1.61 6.86 Base. Ox. #2 (Test B) 10.0 4.39 5.65 9.22E-04 2.42 3.22 Avg. Base. Ox. 10.1 2.98 7.11 9.27E-04 2.01 5.10 Low Air (Test C) 9.89 1.14 8.75 9.34E-04 1.58 7.17 Medium Temp. (Test E) 9.94 5.11 4.83 9.93E-04 2.13 2.70 High Temp. (Test F) 10.1 2.02 8.05 9.94E-04 2.38 5.67  Sample 201 had much more varied cyanide consumption results. The highest levels of consumption were seen for baseline cyanidation, baseline oxidation, low airflow tests and high temperature tests. Sample 201 was the only sample to receive thiocyanate testing for all its PLS samples. Cyanide consumption due to thiocyanate production was highest for medium and high temperature tests, indicating that sulphur dissolved during these oxidation treatments may not have been fully oxidized to sulphate. Compared to samples 167 and 169 however, relatively lower amounts of thiocyanate are being produced for this ore. Cyanide consumption was not shown to significantly affect gold extraction. Test C had a significantly lower level of gold extraction compared to other tests (~72%) but only moderate levels of cyanide consumption; the low gold extraction value for this test may be attributed to the nugget effect, which brought down the total gold extraction with a higher than normal level of gold in the tails residue of this trial. Overall, all the above discussion for each ore must be read with care, as the scatter seen in the cyanide consumption results is quite high. Many possible factors could influence these results, including dilution of 96  the leachate solution, degradation of cyanide over time, particularly oxidation of cyanides or reaction with partially oxidized sulphur that result in higher concentrations of either cyanate of thiocyanate. To summarize the results of cyanide consumption, the following tables have been compiled: Table 22: Effect of Measured Parameters on Total Cyanide Consumption Parameter 165 167 169 201 Caustic Dosage Increases consumption Decreases consumption Oxidation Increases consumption Decreases consumption (201 trend fluctuates) Rate of Air Input Increases consumption Decreases consumption (167, 201 trends fluctuate) Dissolved Oxygen Increases consumption Decreases consumption Increases consumption Temperature Increases consumption (201 trend fluctuates)  Table 23: Effect of Measured Parameters on Thiocyanate Production Parameter 169 201 Caustic Dosage Fluctuates (decreases overall) Increases Thiocyanates Oxidation Decreases Thiocyanates Increases Thiocyanates Rate of Air Input Fluctuates (decreases overall) Increases Thiocyanates Dissolved Oxygen Increases Thiocyanates Decreases Thiocyanates Temperature Decreases Thiocyanates Increases Thiocyanates  Given below are a few common behaviours observed for the four ores tested in this work: - as sulphur oxidation increases during pre-treatment, more thiocyanate is evolved in the gold leaching stage, due to more aqueous sulphur species interacting with cyanide - as temperature increases, cyanide consumption due to cyanide oxidation to cyanate increases - High sulphur ores (167, 169) show reduced cyanide consumption as caustic consumption increases - High sulphur ores show reduced cyanide consumption as rates of air input increase - High sulphur ores show reduced cyanide consumption as oxidation temperatures increase  97  4.3 Results of XRD Phase Analysis of Tails Residues  X-Ray Diffraction (XRD) Phase Analysis was performed on untreated (head) samples of all ores, as well as three samples that were chosen from each series of tests – one from baseline cyanidation (with no oxidation), one from baseline oxidation, and one from the high temperature test (typically the test with the highest extent of oxidation). A brief discussion of the results is given below. Rietveld analysis was not conducted for the tails samples, so percentage values of each detected compound cannot be taken at face value, but rather indicate the relative strengths of each match in the ore. However, a Rietveld analysis was performed on the untreated Head samples for each ore through the Earth and Ocean Sciences department of UBC, and thus for the head samples the percentage values associated with each phase can be trusted. Also, clay and other gangue compounds were not searched for; only the main silica (Quartz) phase was detected, along with phases of interest such as iron oxides, iron hydroxides, and sulphur. While other results have been omitted for the sake of brevity, the diffraction patterns of the untreated samples of ore are included in Appendix C for reference. 4.3.1 Sample 165  The top 6 phases detected in the three residues tested for sample 165 are as follows: Table 24: Results of XRD Analysis for Sample 165 (Excluding Clay Minerals)  Head Sample Baseline Cyanidation Baseline Oxidation High Temperature Oxidation Phase 1 Quartz (SiO2): 90.9% Quartz: 94.8% Quartz: 94.3% Quartz: 96.8% Phase 2 Barite (BaSO4): 6.2% Barite: 2.6% Barite:2.7% Barite: 2.1% Phase 3 Native Iron (Fe): 0.4% Magnetite:1.6% Goethite (FeOOH): 1.9% Goethite: 0.4% Phase 4 Hematite (Fe2O3): 0.3% Iron Oxide Hydroxide (FeOOH): 0.5% Ferrihydrite (Fe(OH)3):0.8% Magnetite: 0.3% Phase 5 Magnetite (Fe3O4): 0.1% Mackinawite (FeS): 0.3% Marcasite (FeS2): 0.4% Hematite (proto): 0.2% Phase 6  Hematite: 0.2% Pyrite (FeS2): 0.1% Pyrite: 0.1% 98   Generally, as the level of oxidation rises, the matched spectra for quartz becomes stronger more iron sulphide minerals in the ore degrade. Native iron was found to be present in the initial sample of ore, possibly due to iron introduced in the comminution process. The presence of compounds such as magnetite and hematite increase as iron oxidation was detected, and iron oxy-hydroxides such as ferrihydrite and goethite form in the alkaline slurry. With an ore like sample 165, which does not have much iron or sulphur content, it is very difficult to accurately determine the oxidation states and precipitates that form in the ore without Rietveld refinement. While the exact percentage values for phases present are hard to determine, these results show that most of the iron oxidized reached a valence of +3, and precipitated as goethite, ferrihydrite, or hematite.  4.3.2 Sample 167  The top 6 phases detected in the three residues tested for sample 167 are as follows: Table 25: Results of XRD Analysis for Sample 167 (Excluding Clay Minerals)  Head Sample Baseline Cyanidation Baseline Oxidation High Temperature Oxidation Phase 1 Quartz (SiO2): 78.0% Quartz: 63.2% Quartz: 65.8% Quartz: 72.2%  Phase 2 Pyrite (FeS2): 5.6% Pyrite: 12.4% Marcasite: 11.0% Marcasite: 11.7% Phase 3 Marcasite (FeS2): 4.5% Iron Oxide Hydroxide (FeOOH): 10.6% Pyrite: 9.6% Pyrite: 7.3% Phase 4 Gypsum (CaSO4·2H2O): 2.4% Marcasite: 9.3% Hematite (proto) (Fe2O3): 6.7% Hematite: 3.0% Phase 5 Barite (BaSO4): 1.8% Magnetite (Fe3O4): 2.3% Goethite (FeOOH): 3.0% Goethite: 2.6% Phase 6  Iron Sulphide (FeS): 1.2% Magnetite: 3.0% Iron Oxide (FeO): 1.9%  For sample 167, the most noticeable phase present in the tails residues is goethite or iron oxide hydroxide, which matches well after baseline cyanidation, but is not as strongly matched after oxidative treatment. Iron oxides such as magnetite, hematite, or wustite are present in small quantities after cyanidation and 99  oxidation. What is interesting to note is that the strengths of the patterns matched for iron sulphide minerals present in the oxidized ore increase dramatically compared with the original untreated head sample. This could be because the presence of clay minerals such as kaolinite, illite or muscovite were not considered in XRD analysis of the tails residue as they were for the head samples. Oxidation product evenly contain tri- and di-valent iron, present as hematite, goethite, magnetite, or wustite. 4.3.3 Sample 169  The top 6 phases detected in the three residues tested for sample 169 are as follows: Table 26: Results of XRD Analysis for Sample 169 (Excluding Clay Minerals)  Head Sample Baseline Cyanidation Baseline Oxidation High Temperature Oxidation Phase 1 Quartz (SiO2): 65.3% Quartz: 95.2% Quartz: 62.6% Quartz – 86.6%  Phase 2 Marcasite (FeS2): 12.7% Marcasite: 2.7% Marcasite: 16.8% Marcasite – 6.3% Phase 3 Pyrite (FeS2): 11.0% Pyrite: 1.6% Pyrite: 12.7% Pyrite – 2.4% Phase 4 Gypsum (CaSO4·2H2O): 3.3% Magnetite (Fe3O4): 0.3% Magnetite-h: 5.3% Iron Hydroxide (Fe(OH)2): 2.1% Phase 5 Jarosite (K2Fe6(SO4)4(OH)12): 0.5% Pyrrhotite (Fe7S8): 0.2% Goethite (FeOOH): 2.2% Goethite: 1.5% Phase 6   Hematite (Fe2O3): 0.3% Iron Oxide (FeO): 1.0%  The diffraction patterns for sample 169 follow a noticeable trend – the first three ranked phases in all samples remain the same: quartz, marcasite, and pyrite. As oxidation levels increase, goethite is formed as expected. What is surprising however, is the presence of iron oxides such as magnetite and wustite, even for residues of highly oxidized ore samples. The increased presence of magnetite, which has a mixture of both ferrous and ferric iron, shows that iron is not being fully oxidized to its ferric state, despite a high observed level of sulphur oxidation (evidenced by significant release of aqueous sulphur into solution). Hematite is formed with baseline oxidation conditions but is not detectable for the high temperature oxidation residue. For high temperature oxidation, ferrous iron hydroxide is matched with a moderate 100  degree of confidence after pyrite. This result implies that most of the iron that was oxidized during the high temperature pre-treatment phase ended up with an oxidation state of +2, rather than an expected +3, as would be the case for iron oxy-hydroxide and goethite. 4.3.4 Sample 201  The top 6 phases detected in the three residues tested for sample 201 are as follows: Table 27: Results of XRD Analysis for Sample 201 (Excluding Clay Minerals)  Head Sample Baseline Cyanidation Baseline Oxidation High Temperature Oxidation Phase 1 Quartz (SiO2): 85.7% Quartz: 97.5% Quartz: 78.8% Quartz: 76.8%  Phase 2 Pyrite (FeS2): 3.1% Pyrite: 1.1% Magnetite (Fe3O4): 9.7% Pyrite: 8.5% Phase 3 Gypsum (CaSO4·2H2O): 2.8% Iron Oxide Hydroxide (FeOOH): 0.6% Pyrite: 6.0% Goethite (FeOOH): 6.8% Phase 4 Barite (BaSO4): 1.7% Schwertmannite (FeO2): 0.3% Marcasite: 2.5% Ferrihydrite (Fe(OH)3): 5.6% Phase 5 Jarosite (K2Fe6(SO4)4(OH)12): 0.7% Hematite (proto) (Fe2O3): 0.3% Mackinawite (FeS): 1.0% Pyrrhotite: 2.4% Phase 6 Marcasite (FeS2): 0.7% Pyrrhotite (Fe7S8): 0.3% Bernalite (Fe(OH)3): 1.0%   Sample 201 shows a variety of different oxidation products. For Baseline Cyanidation tests, the highest phases present were unreacted pyrite, followed by iron oxide hydroxide and iron oxides. Hematite was also present in the tails residue, which, when paired with the oxy-hydroxide phase, indicates that most of the iron in the slurry was oxidized to a ferric (+3) state, even without forced aeration. The results of XRD analysis show that there may be some oxidation occurring without air sparging, possibly due to natural ingress of air from vigorous agitation, or maybe some anaerobic oxidation process occurring at high temperatures with low dissolved oxygen levels. For baseline oxidation, magnetite is seen in relatively substantial quantities compared with other oxidation products, a ferrihydrite-like phase called bernalite is 101  also matched but is present in very minor amounts. For the high temperature oxidation tails residue, goethite and ferrihydrite are matched, along with unreacted pyrite, which is to be expected for all ores.  4.4 Comparison of Ores and Discussion  Chapter two previously laid out various oxidation reactions that iron sulphide ores could undergo at different pH and electrochemical potentials. Iron oxides that could be evolved include wustite (FeO), magnetite (Fe3O4), and hematite (Fe2O3), of which the latter two were generally seen to be the most common “stable” oxidative products in alkaline leach tests. Iron hydroxides that could be formed on the surface of mineral particles include ferrous hydroxide (Fe(OH)2) and ferrihydrite (Fe(OH)3). However, the most commonly found by-product for the tests described above seemed to be goethite (FeOOH) or “Iron Oxide Hydroxide” as termed by the Match! Software used for XRD analysis (in Section 4.3 above). Goethite and other iron oxide hydroxide precipitates were found in most of the cited oxidation studies discussed previously. The general method of pyrite/marcasite oxidation and goethite precipitation in alkaline media is given below: FeS2(s) + 7O2(aq) + 2H2O(l) = 2FeSO4(aq) + 2H2SO4(aq)     (eq. 100a) 4FeSO4(aq) + O2(aq) + 2H2SO4(aq) = 2Fe2(SO4)3(aq) + 2H2O(l)    (eq. 100b) (2)(a) + (1)(b) = c 4FeS2(s) + 15O2(aq) + 2H2O(l) = 2Fe2(SO4)3(aq) + 2H2SO4(aq)     (eq. 100c) H2SO4(aq) + 2NaOH(aq) = Na2SO4(aq) + 2H2O(l)      (eq. 100d) (1)(c) + (2)(d) = e 4FeS2(s) + 15O2(aq) + 4NaOH(aq) = 2Na2SO4(aq) + 2Fe2(SO4)3(aq) + 2H2O(l)   (eq. 100e) Fe2(SO4)3(aq) + 6NaOH(aq) = 3Na2SO4(aq) + 2FeOOH(s) + 2H2O(l)    (eq. 100f) (1)(e) + (2)(f) 4FeS2(s) + 15O2(aq) + 16NaOH(aq) = 4FeOOH(s) + 8Na2SO4(aq) + 6H2O(l)   (eq. 100) Eq. 100 above is considered by most researchers to be stoichiometrically accurate for complete pyrite oxidation in caustic media forming goethite. It should be noted that at high pH, very little iron is left in solution for significant periods of time, but very quickly precipitate to either iron oxides or hydroxides. Simultaneous oxidation of pyrite with oxygen as well as ferrous sulphate (FeSO4) to ferric sulphate (Fe2(SO4)3) must occur to either oxidize pyrite and marcasite, or to precipitate goethite in the manner 102  described above. The short lifespan of aqueous iron in alkaline solution could explain why a lot of iron oxides such as magnetite or wustite are present in the tails residues of these ores. Iron found in sulphides could be left unoxidized, leaving its oxidation state as (+2) instead of oxidized to ferric iron (+3). Ferric iron would be required for the formation of goethite. In cases where oxidation very rapidly occurred (as seen in medium or high temperature tests), significant levels of sulphate were detected, indicating high extents of sulphur oxidation. Leached residues from these tests showed that iron compounds such as hematite and goethite were present, showing that when rapid oxidation occurred, the electrochemical potential in the slurry was high enough that ferric iron was produced, rather than ferrous iron. While elevated levels of caustic consumption are most likely due to an increase in the amount of sulphuric acid released from pyrite oxidation, another explanation is offered through the electrochemical redox reactions given below. As previously mentioned, compounds such as goethite, ferrihydrite, and ferrous iron hydroxide all consume significant amounts of caustic sodium hydroxide and could account for a significant portion of the measured consumption in high sulphur ores (e.g. samples 167, 169). Initial: FeS2(s) = FeOOH(s) + 2SO4-2(aq)       (eq. 101a) Oxidation: FeS2(s) + 10H2O(l) = FeOOH(s) + 2SO4-2(aq) + 19H+ + 15e-   (eq. 101b) Reduction: O2(aq) + 4H+ + 4e- = 2H2O(l)       (eq. 102) Overall Redox: (4)(Oxidation) + (15)(Reduction) 4FeS2(s) + 10H2O(l) + 15O2(aq) = 4FeOOH(s) + 8SO4-2(aq) + 16H+(aq)    (eq. 103a) In a caustic environment (adding equal amounts of hydroxide ions to both sides): 4FeS2(s) + 15O2(aq) + 16OH-(aq) = 4FeOOH(s) + 8SO4-2(aq) + 6H2O(l)    (eq. 103b) Alternatively: 4FeS2(s) + 15O2(aq) + 16OH-(aq) = 4Fe(OH)3(s) + 8SO4-2(aq) + 2H2O(l)   (eq. 104) or 4FeS2(s) + 14O2(aq) + 16OH-(aq) = 4Fe(OH)2(s) + 8SO4-2(aq) + 4H2O(l)   (eq. 105) The above reactions show that four moles of caustic are consumed for every mole of goethite, ferrihydrite, or ferrous iron hydroxide formed. In the same way as shown above, caustic requirements for all possible iron oxides and iron hydroxides can be calculated. Caustic will be consumed for every oxidation product observed in the tails residues described in Section 4.3 above. Three parameters were tracked in this work: caustic consumption, rate of airflow, and oxidation temperature. Of these, caustic consumption was not independently controlled with a specific dosage of sodium hydroxide, but rather could increase as the 103  oxidation of pyrite and marcasite progressed. This choice meant that “caustic consumption” is merely an indirect measure of the amount of both sulphuric acid released from oxidized sulphides, as well as the various oxidation products that were produced as a result. Additionally, caustic was likely responsible for dissolving the ore samples and exposing more iron sulphides to an oxidative environment. For future testing, pH control will be an important variable tested. In this body of research, pH was consistently kept at 11. All the tests conducted for these ores may be repeated with perhaps the same methodology (or one that is slightly updated to reflect the findings in this work), at different alkaline pH ranges. With a clear picture of the behaviour of these ores at various pH values from 7 – 12, a better understanding of alkaline oxidation can be achieved, and hopefully even higher gold extractions can ultimately be realised. Since all the tested ores have a wide range of sulphur content, it is difficult to compare them. A summary of the most relevant results obtained in this work are given in Table 28 below. Note that the values given for potential gold revenue use the gold head grade for each ore obtained through Fire Assay, rather than the calculated gold head grade. An explanation of the economic calculations is given below in Section 4.5. Table 28: Summary of Results for Alkaline Oxidation and Cyanidation  Calculated Sulphur Grade (kg/t) NaOH Consumption (kg/t) Sulphur Oxidation (%) Calculated Gold Head Grade (g/t) Gold Extraction (%) Potential Net Revenue Per Tonne ($ USD) 165 Base. Cyan. #1  0.00  11.1 77.8% $474.96 165 Base Cyan. #2  0.00  12.7 80.9% $494.12 165 Avg. Base. Cyan.  0.00  11.9 79.4% $484.54 165 Base. Ox. #1 (Test A) 11.5 7.51 3.89% 15.5 85.5% $518.08 165 Base. Ox. #2 (Test B) 11.4 7.49 5.59% 15.5 84.2% $510.50 165 Avg. Base. Ox. 11.5 7.50 4.74% 15.5 84.8% $514.29 104   Calculated Sulphur Grade (kg/t) NaOH Consumption (kg/t) Sulphur Oxidation (%) Calculated Gold Head Grade (g/t) Gold Extraction (%) Potential Net Revenue Per Tonne ($ USD) 165 Low Air (Test C) 11.9 7.56 6.92% 12.7 81.7% $495.10 165 High Air (Test D) 11.7 7.55 7.41% 12.6 80.8% $489.60 165 Medium Temp. (Test E) 11.2 7.53 6.21% 13.0 81.0% $491.02 165 High Temp. (Test F) 12.0 7.52 4.65% 13.1 82.5% $499.66 167 Base. Cyan. #1  19.00  18.7 86.5% $711.18 167 Base Cyan. #2  25.12  17.7 86.6% $709.47 167 Avg. Base. Cyan.  22.06  18.2 86.5% $710.32 167 Base. Ox. #1 (Test A) 82.6 28.81 26.4% 18.1 88.2% $721.05 167 Base. Ox. #2 (Test B) 99.0 37.49 40.8% 17.3 90.8% $737.83 167 Avg. Base. Ox. 90.8 33.15 33.6% 17.7 89.5% $729.44 105   Calculated Sulphur Grade (kg/t) NaOH Consumption (kg/t) Sulphur Oxidation (%) Calculated Gold Head Grade (g/t) Gold Extraction (%) Potential Net Revenue Per Tonne ($ USD) 167 Low Air (Test C) 95.4 34.9 38.6% 7.43 90.2% $734.62 167 High Air (Test D) 97.7 38.9 40.1% 7.33 89.6% $727.55 167 Medium Temp. (Test E) 122 40.9 55.6% 18.1 90.9% $737.78 167 High Temp. (Test F) 130 45.2 58.8% 18.5 91.8% $742.41 169 Base. Cyan. #1  16.3  7.16 66.8% $207.53 169 Base Cyan. #2  16.1  7.40 65.0% $201.57 169 Avg. Base. Cyan.  16.2  7.28 65.9% $204.55 169 Base. Ox. #1 (Test A) 175 33.8 36.7% 7.24 72.5% $217.43 169 Base. Ox. #2 (Test B) 182 38.2 37.2% 7.12 72.2% $214.08 169 Avg. Base. Ox. 178 36.0 37.0% 7.18 72.4% $215.76 106   Calculated Sulphur Grade (kg/t) NaOH Consumption (kg/t) Sulphur Oxidation (%) Calculated Gold Head Grade (g/t) Gold Extraction (%) Potential Net Revenue Per Tonne ($ USD) 169 Low Air (Test C) 196 34.8 42.6% 7.43 72.4% $216.64 169 High Air (Test D) 191 43.3 38.2% 7.33 71.9% $210.45 169 Medium Temp. (Test E) 202 34.6 50.0% 7.42 83.0% $214.66 169 High Temp. (Test F) 249 42.7 61.5% 7.42 90.7% $214.82 201 Base. Cyan. #1  14.6  4.17 81.0% $133.49 201 Base Cyan. #2  24.4  4.21 81.1% $128.89 201 Avg. Base. Cyan.  19.5  4.19 81.0% $131.19 201 Base. Ox. #1 (Test A) 42.2 21.4 38.3% 4.26 88.0% $142.63 201 Base. Ox. #2 (Test B) 43.2 27.0 38.7% 4.35 85.3% $134.98 201 Avg. Base. Ox. 42.7 24.2 38.5% 4.30 86.6% $138.81 107   Calculated Sulphur Grade (kg/t) NaOH Consumption (kg/t) Sulphur Oxidation (%) Calculated Gold Head Grade (g/t) Gold Extraction (%) Potential Net Revenue Per Tonne ($ USD) 201 Low Air (Test C) 42.5 21.5 36.8% 5.23 71.8% $113.73 201 High Air (Test D) 39.6 33.7 37.9% 4.31 89.2% $138.64 201 Medium Temp. (Test E) 44.2 35.7 49.4% 4.81 83.0% $126.59 201 High Temp. (Test F) 49.1 32.7 51.% 4.40 90.7% $141.80  A few behaviours that are common to most, if not all the tested ores, will be discussed below. For all plots, a separate series is plotted with all data points from all ores. Additional series comprised of data from individual ores are overlaid on top, indicated below with different markers. Where applicable, an overall trend line is plotted (in dark blue). If the trendline is plotted for only three of the four ores tested, it will be stated in the caption of each plot. Also, where possible, the baseline cyanidation test results (corresponding to no oxidation) will be plotted. 108    Figure 57: Gold extraction vs. extent of oxidation for tested ores; baseline cyanidation results included Except for ore 165, the iron-sulphide ores tested responded favourably to sulphur oxidation. When oxidized, sample 165 shows decreasing gold extraction, although all gold leaching results after oxidation were more positive than those tests with just baseline cyanidation stages. Sample 167 shows an almost linear correlation between extent of oxidation and gold extraction. In contrast, despite a significant increase between baseline cyanidation trials and tests with oxidation, the gold extraction results for sample 169 appear to plateau with increasing oxidation. This could be due to an additional refractory concern such as silicate occlusion of gold, or carbonaceous material that may remove gold cyanide complexes from the leachate. Sample 201 also shows a significant increase in gold extraction after oxidation. Given below are plots of extent of oxidation and gold extraction vs. caustic consumption for all tested ores. 65%70%75%80%85%90%95%0% 10% 20% 30% 40% 50% 60% 70%Gold Extraction (%)Extent of Oxidation (%)Au165167169201109    Figure 58: Extent of oxidation vs. total caustic consumed for tested ores; baseline cyanidation results included Again, all tested ores respond favourably with caustic addition. The likeliest explanation for this behaviour is that caustic sodium hydroxide increased as a direct result of increased sulphur oxidation releasing sulphuric acid into solution, which was automatically neutralized by caustic addition. Caustic addition could have also acted to attack gangue elements of the ore, exposing more iron sulphides to the oxidizing environment.   Figure 59: Gold extraction vs. total caustic consumed for tested ores; baseline cyanidation results included (plotted on x=0 for sample 165, shown as the data with the lowest caustic dosages for ores 167 and 169, and at 14.6 kg/t and 24.4 kg/t caustic dosage for sample 201, respectively) 0%10%20%30%40%50%60%70%0 10 20 30 40 50Extent of Oxidation (%)Total Caustic Dosed (kg/t)Ox16516716920160%65%70%75%80%85%90%95%0 10 20 30 40 50Gold Extraction (%)Total Caustic Dosed (kg/t)Au165167169201110  Additionally, increased caustic consumption may indicate that goethite (FeOOH) is being formed as an oxidation product due to the destruction of iron sulphides in these ores (oxidizing iron to a ferric state). Given below in Figure 60 and Figure 61 are plots of oxidation, and gold extraction vs. rate of air input, respectively.   Figure 60: Extent of oxidation vs. rate of air input at 0.68 atm (10 psi) pressure for tested ores; alkaline oxidation results only    Figure 61: Gold extraction vs. rate of air input at 0.68 atm (10 psi) pressure for tested ores; baseline cyanidation results included 0%5%10%15%20%25%30%35%40%45%1 1.5 2 2.5 3Extent of Oxidation (%)Rate of Air Input at 0.68 atm (10 psi) Pressure (LPM)Ox16516716920160%65%70%75%80%85%90%95%0 0.5 1 1.5 2 2.5 3Gold Extraction (%)Rate of Air Input at 0.68 atm (10 psi) Pressure (LPM)Au165167169201111  As shown above, both oxidation and gold extractions increase with all ores with the addition of an aerated pre-treatment stage. Further testing must be performed to indicate whether this trend continues as air sparging rates increase, or whether some limit exists that will result in the highest possible gold extraction for a mixture of all ores. The problem will then be an optimization issue, depending on the ratio of each ore leached at any one time. The low extraction seen at 1.5 LPM for ore 201 can be attributed to the “nugget” effect (gold particle sparsity effect) causing a larger than normal gold concentration in the tails residue, thus lowering the overall extraction of gold seen for this test. The plot of extent of oxidation vs. dissolved oxygen in solution during pre-treatment is given below:  Figure 62: Extent of oxidation vs. dissolved oxygen in solution during oxidation for tested ores; alkaline oxidation results only; trendline plotted for all ores except sample 165 For the above plot, a distinct pattern emerges for the data of ores 167, 169, and 201. Oxidation appears to decrease with the average dissolved oxygen in solution during pre-treatment. Since most these ores responded favourably to high temperature conditions, and high temperatures reduce the amount of oxygen in solution, it may be that this trend is merely a correlation between favourable oxidation conditions and recorded dissolved oxygen data. The trend seen above could also be due to increased oxidation during these tests uses more oxygen in solution than is being added through air sparging or natural uptake, giving a lower average value of oxygen in solution throughout those tests. Given below in Figure 63 and Figure 64 are plots of extent of oxidation and gold extractions vs. oxidation temperature. 0%10%20%30%40%50%60%70%0 1 2 3 4 5 6 7Extent of Oxidation (%)Dissolved Oxygen in Solution During Oxidation (mg/L)Ox165167169201112   Figure 63: Extent of oxidation vs. oxidation temperature for tested ores; alkaline oxidation results only  Figure 64: Gold extraction vs. oxidation temperature for tested ores; alkaline oxidation results only The two plots above show strong trends for increased oxidation as well as gold extractions for all ores. High sulphide ores especially show a marked improvement for sulphur dissolution. However, ore 165 shows no significant improvement with oxidation temperature. This is likely due to the extremely low sulphur content for sample 165, and so sulphur oxidation may not impact gold extractions like with the other samples. Ore 169 also shows no significant change in gold extraction as temperatures rise. The trendline for the gold extraction graph uses data from all tested ores. One other parameter that may be relevant to this discussion is electrochemical potential. The relevant plot is given below: 0%10%20%30%40%50%60%70%30 35 40 45 50 55 60Extent of Oxidation (%)Oxidation Temperature (°C)Ox16516716920165%70%75%80%85%90%95%30 35 40 45 50 55 60Gold Extraction(%)Oxidation Temperature (°C)Au165167169201113   Figure 65: Extent of oxidation vs. average electrochemical potential observed during oxidation for tested ores; alkaline oxidation results only; trendline plotted for all ores except sample 165 All ores except sample 165 show decreasing extent of oxidation with increased electrochemical potential in the slurry. The above potentials are the averaged values of potentials over the course of each trial, and do not show how potential generally rose with time throughout oxidation. In cases where significant amounts of caustic were introduced to the slurry, potentials hovered around 230 mV SHE while caustic sodium hydroxide was being introduced to the slurry. Potentials then increased with time as the pH of the slurry dropped (within tolerance – caustic addition was automatically triggered at 10.90 pH). What is interesting to note is that the high and medium temperature tests for ores 167, 169, and 201 resulted in the highest levels of oxidation but also had the lowest electrochemical potentials in the slurry. It is not clear why this is, as generally the higher potentials observed in the slurry, the higher the electrochemical drive is to oxidize the sulphide ore.  4.5 Economic Analysis of Test Results  So far, a few qualitative trends have been seen in different ores, but the concrete benefit of pre-oxidative treatment of the ores tested for this work has yet to be quantified. In this section, the economic benefit of processing these ores prior to the gold leaching step is explained. In all cases, the introduction of forced air sparging and addition of caustic NaOH improves the net revenue that could potentially be gained per tonne of ore processed, except when a “nugget effect” lowers the total amount of gold leached due to increased gold detected in the tails residue of a trial. To calculate these 0%10%20%30%40%50%60%70%150 170 190 210 230 250 270 290Extent of Oxidation (%)Electrochemical Potential During Oxidation (mV SHE)Ox165167169201114  values, only the cost of reagent addition was compared with the increase of gold extracted per tonne of ore. The cost of air, or energy required for heating or pumping equipment was not factored in to the calculation. For the gold grade for each ore, the basis chosen for all trials was the averaged value obtained through Fire Assay tests of the untreated “head” samples of each ore. The industry average was used for the price of caustic sodium hydroxide, and the price of gold was taken at the beginning of February 2018. They are as follows: NaOH: $0.50/kg USD Gold: $43.25/g USD Given below in Figure 66 are the net revenues per tonne of ore processed that would be achieved with all ore samples tested in this work.  Figure 66: Potential Achievable Net Revenue Per Tonne Ore for All Ores Tested The net revenues for different ores are extremely varied. It is necessary to briefly discuss each ore separately. The cost-benefit breakdown for each ore is given below.$0$100$200$300$400$500$600$700$800B1 B2 O1 O2 C D E FNet Revenue per Tonne Ore Processed($ USD)Test Performed165167169201115  4.5.1 Sample 165  The economic calculations for ore 165 (gold head grade: 14.16 g/mt) are shown below in Table 29. Table 29: Summary of Economic Calculations for Sample 165 Test Gold Ext. (%) Increased Gold Ext. (%) Increased Gold Ext. per Tonne Ore (g/mt) Increased Gold Revenue ($USD) NaOH Dosed (kg/t) NaOH Cost ($USD) Net Rev. ($USD) Net Rev. Benefit ($USD) Base. Cyan. #1 77.8%    0.00 $0.00 $476.66  Base Cyan. #2 80.9%    0.00 $0.00 $495.82  Avg. Base. Cyan. 79.5%    0.00 $0.00 $486.81  Base. Ox. #1 (Test A) 85.5% 6.00% 0.85 $36.73 7.51 $3.76 $519.78 $32.97 Base. Ox. #2 (Test B) 84.2% 4.76% 0.67 $29.13 7.49 $3.75 $512.20 $25.39 Avg. Base. Ox. 84.8% 5.38% 0.76 $32.94 7.50 $3.75 $515.99 $29.19 Low Air (Test C) 81.7% 2.25% 0.32 $13.77 7.56 $3.78 $496.80 $9.99 High Air (Test D) 80.8% 1.35% 0.19 $8.26 7.55 $3.77 $491.30 $4.49 116  Test Gold Ext. (%) Increased Gold Ext. (%) Increased Gold Ext. per Tonne Ore (g/mt) Increased Gold Revenue ($USD) NaOH Dosed (kg/t) NaOH Cost ($USD) Net Rev. ($USD) Net Rev. Benefit ($USD) Medium Temp. (Test E) 81.0% 1.58% 0.22 $9.68 7.53 $3.77 $492.72 $5.91 High Temp. (Test F) 82.5% 2.99% 0.42 $18.31 7.52 $3.76 $501.36 $14.55  The highest gold extractions for sample 165 were achieved using baseline oxidation conditions. On average, baseline oxidation would increase the net revenue per tonne of ore processed by $29.19. Due to the low sulphur grade of sample 165 no caustic was needed during oxidation to keep pH at 11, since the lime used during grinding may have already contributed to sulphur oxidation prior to these tests. All oxidation test conditions resulted in an increase in potential net revenue from the conventional cyanidation. Extraction were generally quite high for sample 165, making this one of the most economically favourable ores tested. High temperature oxidation also increased the net revenue for sample 165, a result which is repeated with subsequent ore samples. For sample 165 and indeed all samples tested, more repetitions may be needed to confirm these results, especially for those test conditions which did not have repeats to average, unlike baseline cyanidation and baseline oxidation tests.        117  4.5.2 Sample 167  A summary of the economic calculations for ore 167 (gold head grade: 16.73 g/mt) is given below. Table 30: Summary of Economic Calculations for Sample 167 Test Gold Ext. (%) Increased Gold Ext. (%) Increased Gold Ext. per Tonne Ore (g/mt) Increased Gold Revenue ($USD) NaOH Dosed (kg/t) NaOH Cost ($USD) Net Rev. ($USD) Net Rev. Benefit ($USD) Base. Cyan. #1 86.5%    19.0 $9.50 $714.18  Base Cyan. #2 86.6%    25.1 $12.56 $712.47  Avg. Base. Cyan. 86.6%    22.1 $11.03 $713.49  Base. Ox. #1 (Test A) 88.2% 1.67% 0.32 $13.94 28.8 $14.41 $724.05 $10.56 Base. Ox. #2 (Test B) 90.8% 4.19% 0.81 $35.05 37.5 $18.75 $740.83 $27.34 Avg. Base. Ox. 89.5% 2.90% 0.56 $24.28 33.2 $16.58 $732.23 $18.74 Low Air (Test C) 90.2% 3.65% 0.71 $30.53 34.9 $17.43 $737.62 $24.13 High Air (Test D) 89.6% 3.04% 0.59 $25.48 38.9 $19.45 $730.55 $17.06 118  Test Gold Ext. (%) Increased Gold Ext. (%) Increased Gold Ext. per Tonne Ore (g/mt) Increased Gold Revenue ($USD) NaOH Dosed (kg/t) NaOH Cost ($USD) Net Rev. ($USD) Net Rev. Benefit ($USD) Medium Temp. (Test E) 90.9% 4.38% 0.85 $36.69 40.9 $20.42 $740.78 $27.30 High Temp. (Test F) 91.8% 5.20% 1.01 $43.49 45.2 $22.60 $745.41 $31.92  Sample 167 responded extremely well to oxidation treatment. The highest potential revenues per tonne of ore were those stemming from high and medium temperature tests – although the combination of increased oxidation as well as lower sodium hydroxide requirements during baseline oxidation allow net revenues from those tests to also reach high levels. Although it required significant amounts of caustic to dissolve ore samples, the high gold extraction levels achieved during tests indicates that this ore is very economically viable to process, despite high reagent costs. Elevated temperatures increased the net revenue benefit that could be achieved with sample 167, a favourable result for this type of pre-treatment. For the highest temperature tested, more than 1 g/mt of additional gold could be extracted from the ore, which is a significant amount for precious metals.119  4.5.3 Sample 169  The economic data for ore 169 (gold head grade: 7.57 g/mt) are given below in Table 31. Table 31: Summary of Economic Calculations for Sample 169 Test Gold Ext. (%) Increased Gold Ext. (%) Increased Gold Ext. per Tonne Ore (g/mt) Increased Gold Revenue ($USD) NaOH Dosed (kg/t) NaOH Cost ($USD) Net Rev. ($USD) Net Rev. Benefit ($USD) Base. Cyan. #1 66.8%    16.3 $8.16 $210.53  Base Cyan. #2 64.8%    16.1 $8.05 $204.57  Avg. Base. Cyan. 65.9%    16.2 $8.11 $207.50  Base. Ox. #1 (Test A) 72.5% 6.64% 0.50 $21.72 33.8 $16.90 $220.43 $12.93 Base. Ox. #2 (Test B) 72.2% 6.29% 0.48 $20.58 38.2 $19.10 $217.08 $9.58 Avg. Base. Ox. 72.4% 6.47% 0.49 $21.19 36.0 $18.00 $218.79 $11.29 Low Air (Test C) 72.4% 6.56% 0.50 $21.45 34.8 $17.42 $219.64 $12.14 High Air (Test D) 71.9% 5.96% 0.45 $19.51 43.3 $21.67 $213.45 $5.95 120  Test Gold Ext. (%) Increased Gold Ext. (%) Increased Gold Ext. per Tonne Ore (g/mt) Increased Gold Revenue ($USD) NaOH Dosed (kg/t) NaOH Cost ($USD) Net Rev. ($USD) Net Rev. Benefit ($USD) Medium Temp. (Test E) 71.8% 5.91% 0.45 $19.34 34.6 $17.29 $217.66 $10.16 High Temp. (Test F) 73.1% 7.19% 0.54 $23.54 42.7 $21.32 $217.82 $10.32  In contrast to sample 167, sample 169 was one of the least economically viable ores tested. Poor gold extractions and a low gold grade combined with huge caustic dosage requirements meant that any possible revenues that could be gained from gold leaching were significantly impacted per tonne of ore processed. In some cases, almost a tenth of gold revenue may be lost due to increased caustic dosages, a significant amount, although this might be a normal margin that commercial operators tolerate for lean gold ores. While high temperature oxidation resulted in the highest levels of gold extraction, very high caustic requirements during oxidation meant that net revenues were not as high as baseline oxidation, or oxidation using a low air-flow rate. Overall, baseline oxidation elevated Net Economic Benefit an average of $11.29, and the low airflow oxidation test increased the calculated net revenue benefit by $12.14 per tonne of ore, compared to only $10.16 and $10.32 for medium and high temperature oxidation, respectively. Elevated temperature tests required more than 30 – 40 kg/t of NaOH to maintain a pH of 11 throughout pre-treatment, pH conditioning, and cyanidation, so more than $15 – 20.00 of the gold revenue would be lost just from the cost of caustic reagent. In contrast, only 16.2 kg/t of caustic was needed for baseline cyanidation. This low caustic dosage shows that NaOH may indeed work to dissolve the ore throughout a pre-treatment phase, exposing more sulphur to oxidative conditions, thus releasing more sulphuric acid into the bulk solution – which in turn causes more caustic to be dosed to the leaching reactor. Additionally, a low consumption of caustic may indicate that less ferric iron is being evolved (from oxidation of pyrite and marcasite). This in turn leads to less hydroxide being consumed to form goethite (FeOOH).121  4.5.4 Sample 201  A summary of economic analysis for ore sample 201 (gold head grade: 4.11 g/mt) is given below. Table 32: Summary of Economic Calculations for Sample 201 Test Gold Ext. (%) Increased Gold Ext. (%) Increased Gold Ext. per Tonne Ore (g/mt) Increased Gold Revenue ($USD) NaOH Dosed (kg/t) NaOH Cost ($USD) Net Rev. ($USD) Net Rev. Benefit ($USD) Base. Cyan. #1 81.0%    14.6 $7.30 $136.49  Base Cyan. #2 81.1%    24.4 $12.18 $131.89  Avg. Base. Cyan. 81.0%    19.5 $9.74 $134.21  Base. Ox. #1 (Test A) 88.0% 6.97% 0.29 $12.38 21.4 $10.69 $145.63 $11.42 Base. Ox. #2 (Test B) 85.3% 4.23% 0.17 $7.52 27.0 $13.48 $137.98 $3.77 Avg. Base. Ox. 86.6% 5.58% 0.23 $9.91 24.2 $12.09 $141.77 $7.56 Low Air (Test C) 71.8% -9.28% -0.38 -$16.48 21.5 $10.74 $116.73 -$17.48 High Air (Test D) 89.2% 8.19% 0.34 $14.55 33.7 $16.87 $141.64 $7.42 122  Test Gold Ext. (%) Increased Gold Ext. (%) Increased Gold Ext. per Tonne Ore (g/mt) Increased Gold Revenue ($USD) NaOH Dosed (kg/t) NaOH Cost ($USD) Net Rev. ($USD) Net Rev. Benefit ($USD) Medium Temp. (Test E) 83.0% 1.97% 0.08 $3.50 35.7 $17.86 $129.59 -$4.62 High Temp. (Test F) 90.7% 9.69% 0.40 $17.21 32.7 $16.37 $144.80 $10.59  For sample 201, the highest levels of gold extractions are once again seen after high temperature oxidation. Although the gold revenues that could be achieved under those conditions are quite high, they can almost be matched using baseline oxidation. As previously discussed the tails residue for Test C likely had an occluded gold grain present inside, which brought down the total gold extraction associated with that test. Further tests would be required to confirm that varying airflow in the oxidation phase works to decrease or increase oxidation and subsequent levels of gold extraction for this ore. An interesting result is that medium temperature tests, while increasing gold extractions slightly, requires significant amounts (over 35 kg/t) of caustic to counteract the release of sulphuric acid from sulphur oxidation and formation of goethite, and so would result in a net loss of revenue compared to baseline cyanidation.  4.6 Reproducibility of Tests  As described above, both baseline cyanidation tests as well as baseline oxidation tests were conducted in duplicate to ensure the averaged results of these tests could be considered accurate. These results could then form the basis of comparison with other tests. To compare the reproducibility of these tests, it is important to look at Percent Difference. Percent difference is a measure of how different values are in two different tests compared to the averaged value of those test results. Percent difference is defined in the following equation below: % 𝐷𝑖𝑓𝑓𝑒𝑟𝑒𝑛𝑐𝑒 =  𝐴𝑏𝑠𝑜𝑙𝑢𝑡𝑒 𝑣𝑎𝑙𝑢𝑒 𝑜𝑓 𝑡ℎ𝑒 𝑑𝑖𝑓𝑓𝑒𝑟𝑒𝑛𝑐𝑒 𝑖𝑛 𝑣𝑎𝑙𝑢𝑒𝑠𝐴𝑣𝑒𝑟𝑎𝑔𝑒 𝑜𝑓 𝑡ℎ𝑒 𝑣𝑎𝑙𝑢𝑒𝑠 𝑥 100% =  2 𝑥 |𝑣𝑎𝑙𝑢𝑒 1−𝑣𝑎𝑙𝑢𝑒 2| (𝑣𝑎𝑙𝑢𝑒 1+𝑣𝑎𝑙𝑢𝑒 2) 𝑥 100%   123  The absolute difference between test data can vary greatly depending on what variable is being examined, and if the experiment conditions were under good control at the time. However, with natural ores such as those examined in this work, numerous gangue components present in the ore can alter test results with little or no discernable explanation, despite care being taken to homogenize ore samples.  Percent difference can indicate whether experimental error is too high to make any meaningful conclusions from the data, and whether the data itself can be trusted to be averaged and compared with other results. For gold extraction and oxidation, the most important error is the percent difference between the final values for each set of tests for each ore. As an example, percent differences were calculated comparing final (24-hour) gold extraction levels for Baseline Cyanidation tests #1 and #2 of ore 169, or final (8-hour) oxidation extents for Baseline Oxidation tests #1 and #2 conducted with ore 201. Gold and sulphur concentrations may vary slightly between tests, and so extraction curves may not conform to a smooth, regular shape because of incidental dilution or evaporation, so final extraction values are used. Results of all error calculations are given below in Table 33, Table 34, and Table 35. Table 33: Average Percent Difference Results for Recorded Data for Baseline Cyanidation Tests Stage Physical Parameter Sample 165 Sample 167 Sample 169 Sample 201 Average Percent difference Cyanidation pH 1.13% 0.76% 2.80% 1.58% 1.57% ORP (Eh) 54.1% 13.0% 6.52% 20.1% 23.4% D.O. 2.32% 3.76% 5.71% 4.14% 3.98% Temperature 0.52% 0.24% 0.47% 0.24% 0.37%  Table 34: Average Percent Difference Results for Recorded Data for Baseline Oxidation Tests Stage Physical Parameter Sample 165 Sample 167 Sample 169 Sample 201 Average Percent difference Oxidation pH 0.47% 0.40% 0.71% 0.63% 0.55% ORP (Eh) 5.24% 10.56% 10.35% 10.06% 9.05% D.O. 7.22% 54.1% 14.5% 10.8% 21.8% Temperature 1.32% 0.28% 0.11% 0.64% 0.59% Cyanidation pH 0.86% 1.82% 0.36% 3.65% 1.67% ORP (Eh) 9.43% 2.29% 4.96% 4.36% 5.26% D.O. 6.02% 2.85% 5.41% 5.23% 4.88% Temperature 2.40% 0.28% 0.19% 0.24% 0.78%  124  Table 35: Percent and Absolute Difference of Final Extraction Values from Baseline Tests Tests Stage Difference Sample 165 Sample 167 Sample 169 Sample 201 Average Difference Baseline Cyanidation Cyanidation Percent 3.94% 0.19% 2.82% 0.19% 1.78% Absolute 3.13% 0.16% 1.86% 0.16% 1.33% Baseline Oxidation  Oxidation Percent 35.7% 42.7% 1.37% 0.95% 20.2% Absolute 1.69% 14.3% 0.51% 0.37% 4.23% Cyanidation Percent 1.46% 2.82% 0.48% 3.16% 1.98% Absolute 1.24% 2.52% 0.35% 2.74% 1.71%  Aqueous gold in solution was tested using AAS performed by Inspectorate International Ltd. (Bureau Veritas, S.A.). This technique had an error of 0.05 mg/L. In most cases, this error could vary calculated gold head grades by less than 0.50 g/mt, and final gold extractions by less than 1%. Gold fire assay results were more accurate, to within 0.01 g/mt. For LECO sulphur tests, the error in sulphur grades was 0.02%. This error could potentially vary calculated sulphur grades by 0.2 kg/t (0.2%), and final sulphur oxidation levels by 0.1%. For ICP-MS analysis, had an error of 0.01 – 5 ppm, depending on the element tested. ICP-AES/OES tests generally had an error of 0.1 mg/L. Thiocyanate test results were accurate within 1 mg/L. The error inherent in results using these techniques were not significant compared to the error associated with varying gold, sulphur, and iron grades in the ores, the dilution and evaporation experienced during oxidation and cyanidation tests, and other sources of error, which are discussed below. It is clear from the tables above that the cyanidation results of all tests have the lowest percent difference between test values. This observation holds for both the Baseline Cyanidation tests as well as the cyanidation stages in Baseline Oxidation tests. In addition to the final extraction values being quite close together for all cyanidation tests, the physical data recorded during the cyanidation stages more closely agreed with each other for all tests performed. Examining Table 33 and Table 34, the easiest physical parameter to control over all the tests was temperature, followed by pH, especially in the case of ores 167, 169, and 201, which needed automated caustic dosage to control pH. Baseline oxidation results would have had a very low percent difference between trials, if not for the varied results in ores 165 and 167. Sample 165 had an extremely low assayed sulphur grade (0.88%) so differences between final oxidation levels were extremely sensitive to even slight changes in sulphur concentration. While the absolute difference in sulphur released into solution in each test was quite small, the percent 125  difference between them was significant. It is possible that the sulphur in the test with lower oxidation was occluded from the leachate by a clay or silica coating in the ore. However, since the absolute difference between oxidation levels for these tests is quite low, the true baseline level of oxidation can be taken as the average of the two observed extents of oxidation, with a possible further set of oxidation tests to be performed at a future time. For ore 167, it is unclear what caused the major difference between tests A and B (baseline oxidation tests). Examining the assayed sulphur results of the residues, it seems that there is no discernible trend in how sulphur is being released into solution over time. Test A had LECO results of 6.16% sulphur grade in the solids at hour 8 (end of the aerated oxidation stage), 6.69% at hour 24 (just prior beginning of the pH conditioning and cyanidation stage), and 6.43% in the tails after the cyanidation stage. Test B had 5.91%, 5.78%, and 6.49%, respectively. It may be that the test with lower oxidation results (Test A) may have had a large chunk of unoxidized sulphur occluded from the oxidative conditions in the slurry by a clay or silicate layer, which brought down total oxidation from the average value. However, this explanation is purely speculative, since no surface or product layer analysis was conducted due to time constraints. By far the biggest source of “error” in the physical data was the Oxidation/Reduction or Electrochemical Potential (ORP, or Eh) recorded in the slurries. Vigorous agitation was employed throughout all tests, and all reactors used for oxidation and leaching had inbuilt baffles. At such high stirring speeds, the conditions inside the reactor were assumed to be uniform throughout. Readings depend mainly upon the composition of the slurries themselves, as well as the density of particles in the slurry. A concrete example of this is when the ORP probe would contact a larger sized grain of ore, or when a pebble would become lodged in the probe opening, continuously touching the probe’s platinum wire. In this case, the electrochemical potential reading would be extremely negative, despite the oxidative conditions in the slurry. This reading would indicate a very different result than what is representative of the slurry and is thus inaccurate. It would also create a huge percent difference between test results, since it is unlikely that a similar situation would cause another probe to read a close value in the other reactor at the same time. The recorded Dissolved Oxygen (D.O.) values for all tests were also moderately varied. This is mainly a result of inconsistent air intake experienced inside the reactors, especially during oxidation when air sparging was being performed. As previously discussed, it was difficult to accurately control air sparging rates (and oxygen uptake into solution) because the glass frits used to pump air into the slurries kept getting clogged by small ore particles (<50 μm). Additionally, ensuring complete mixing in the reactors and 126  uniform conditions throughout all oxidation and gold leaching tests may result in a more consistent reading of dissolved oxygen both in terms of probe location, as well as at any given time during each test. Overall, the reproducibility of tests is of acceptable quality, with nearly all test results and data close to each other, as evidenced by percent difference values calculated. The only exceptions are the baseline oxidation values for samples 165 and 167. For physical data recorded, electrochemical potential and dissolved oxygen were seen to be the most varied between two trials under the same conditions, and some explanations for this were given. Despite some results giving high percent differences, it is still possible to average these results and draw conclusions and comparisons based on these values. For future study, more repetitions are needed to confirm the results presented in this work. As repetitions of tests are performed under the same conditions, the results of extent of oxidation, gold extraction, as well as the physical data recorded for each test will trend towards their respective true values. Furthermore, as more fire assay tests and LECO sulphur tests are performed on each of the ores from the White Mountain mine, the averaged head gold and sulphur grades will become increasingly more representative of each ore. It is clear from the previous discussion that the most important factors in obtaining good results are: - proper homogenization of the sample ore samples - good control over all physical parameters, including pH, temperature, caustic addition, and particularly air sparging rates - well adjusted and calibrated recording equipment before every test is undertaken - performing the tests in the same exact way every time - cutting down dilution and evaporation of leachate as much as possible, given washes of probes and reactor walls need to occur at regular intervals  4.7 Summary of Results  A few key results were found through alkaline oxidation and cyanidation of iron sulphide ores. They are summarized below: - the four iron sulphide ores tested achieved superior gold extraction results with the addition of an oxidative pre-treatment phase before gold cyanidation 127  - samples showed a correlation between higher caustic dosage and increased iron sulphide oxidation and subsequently higher gold extractions. Higher levels of caustic consumption may have helped to expose more iron sulphide to an oxidative environment, and more gold to the cyanide leaching medium. Sodium hydroxide (NaOH) consumption increased as more goethite was formed, and more sulphuric acid was released from oxidation of iron sulphides, requiring pH control through caustic addition - ores responded favourably to increasing rates air sparging, which achieved a maximum level of oxidation and gold extraction around the 1.5 – 2.0 LPM mark, with 0.68 atm (10 psi) air pressure for ore 169, and between 2.0 – 2.5 LPM mark for ores 167 and 201. - all samples achieved extremely high levels of sulphur and iron oxidation at high temperatures compared with baseline oxidation, resulting in elevated levels of gold extraction for all ores except sample 165 and 169. More testing may be performed on these ores to determine why gold extraction did not increase, especially as sample 169 experienced a significant increase in sulphide oxidation with higher oxidation temperatures - goethite, magnetite, and hematite, as well as other iron oxides and -hydroxides were produced through iron sulphide oxidation. As the extent of oxidation increased for ores, more iron was found to be oxidized to a ferric (+3) state which precipitated almost immediately as the surface layer products. As oxidation of sulphur increased, more sulphate was released into the leach slurry. Partially oxidized sulphur may have been present as other aqueous compounds in solution such as thiosulphate or sulphite, although this hypothesis may be tested in future through other analyses on solution samples - samples tested offered good gold extraction above 80% with alkaline pre-treatment, except for sample 169, which may have a secondary refractory issue which prevents it from being leached properly with cyanide. Sample 165 offered higher extraction values, but compared to the rest of the tested samples, this ore did not improve as significantly with the addition of an aerated oxidation phase at any rate of air flow, or oxidation temperature, and did not require much significant amounts of caustic sodium hydroxide as oxidation occurred (due to its low sulphur content). - cyanide oxidation and degradation were extremely prevalent during all cyanidation trials. Generally, more cyanate was produced over the course of each trial. However, thiocyanate was evolved because of increased presence of sulphur in the ore, or because of higher concentrations of partially oxidized sulphur compounds such as aqueous thiosulphate or sulphite 128  - are economically favourable to leach, considering only the chemical reagent cost necessary to grind and maintain high pH for the oxidation and cyanidation leach stages. Net revenues for all ores increased with the addition of a pre-treatment phase in nearly all conditions, despite the additional cost associated with dosing caustic to the leach to maintain high pH. The average baseline cyanidation revenues, and highest achievable net revenues per tonne of each ore were: o 165 - $486.81 (baseline cyanidation), $515.99 (baseline oxidation conditions) o 167 - $713.49 (baseline cyanidation), $745.41 (high temperature conditions) o 169 - $207.50 (baseline cyanidation), $219.64 (low rate air-sparging conditions) o 201 - $134.21 (baseline cyanidation), $141.80 (high temperature conditions) - A few ores may exhibit the particle sparsity effect (nugget effect) discussed in previous sections. Ores 165 and 201 seem to have varying calculated gold head grades that have a moderate amount of error, indicating a possible nugget effect. - Ore 169 had the highest calculated sulphur grade compared to its original assay value for sulphur using LECO analysis. A simple explanation for this is that ore sample 169 was very moist after the grinding process, despite extensive pressure filtration. This excess moisture may have contained sulphuric acid evolved during the grinding phase, and this artificially inflated the perceived sulphur grade of the ore used for oxidation trials, particularly those near the end of the test series. The original assay value of sulphur was obtained from a sample that was taken from mainly the top most part of the ore storage container, and this assay sample may have contained less moisture (containing sulphuric acid) than the bulk of the ground ore. This pattern of increasing sulphur grade with subsequent ore samples is evident for the high sulphur ores 167 and 169, shown in Table 28.129  Chapter 5: Conclusions and Recommendations 5.1 Conclusions  In this work, alkaline oxidation pre-treatment was carried out on refractory iron sulphide ores from the White Mountain Mine in the Jilin province in China. Three major parameters were examined in this work: caustic consumption necessary to maintain pH over the course of the tests, the effects of different rates of air sparging at constant temperature, and the temperature of the ore slurries throughout oxidation. Cyanide consumption was also calculated for each leach test, and effects of oxidation were described. X-Ray Diffraction (XRD) analysis was conducted on the tails residues of some of the tests to determine what phases are present at the end of the leach. Economic analysis was conducted on each ore tested to determine whether oxidative pre-treatment increases the financial benefit of processing the ores in this manner. Some conclusions drawn from this work are as follows: The result of baseline cyanidation tests performed on each ore yielded an average value of gold extraction for each ore without prior treatment. These values are as follows: sample 165 – 79.5%, 167 – 86.6%, 169 – 65.9 %, 201 – 81.0%. Trials conducted with prior alkaline oxidative pre-treatment generally yielded higher gold extractions for all conditions tested. Tests with pre-treatment performed under baseline oxidation conditions yielded the following results: 165 – 84.8%, 167 – 89.5%, 169 – 72.4%, 201 – 86.6%. On average, these results were 5.08% better than the baseline cyanidation results. Based on the results baseline cyanidation and baseline oxidation tests, which were obtained by performing each set of tests in duplicate, the reproducibility of the gold leaching tests described in this work is high. Each ore sample had two baseline cyanidation tests and two baseline oxidation tests, with an average percent difference of 1.78% and 1.98% for each set of cyanidation results, respectively. However, the percent difference seen of extent of oxidation values were quite high in some cases. In one case, the large percent difference between oxidation values can be explained by a low sulphur grade which can dramatically influence results with very tiny fluctuations of sulphur in solution or in tails residues. Caustic addition to leach slurries was seen to have a strong correlation to increased pyrite and marcasite oxidation, due to increased release of aqueous sulphur into solution, presumably through generation of sulphuric acid. This in turn required higher levels of caustic dosage to control pH. Caustic consumption was also indicative of iron hydroxides being formed in the tails as oxidation of iron sulphides occurred; mineral such as goethite (FeOOH) and ferrihydrite (Fe(OH)3) may have consumed significant amounts of 130  hydroxide as oxidation progressed for tested samples. Additionally, increased caustic may have led to exposure of iron sulphides to the oxidative environment, by attacking gangue components in the ores. At the pH tested, pre-aeration of ore slurries led to higher extents of oxidation and gold extractions in all ores. Ore samples 167 and 201 responded extremely favourably to increased caustic dosages in the leach in terms of oxidation as well as gold extraction.  Elevated oxidation levels were observed at increasing rates of air sparging. While the curvature of trendlines differed between ores, the general trend was that oxidation tests carried out with forced aeration of the slurry resulted in higher oxidation levels, and subsequently higher gold extractions. Samples 165, 167, and 169 yielded their highest gold extractions (at baseline temperatures of 35°C) with oxidation conducted with 2.0 LPM air sparging. Sample 201 however, yielded higher gold extractions after oxidation with 2.5 LPM air flow. Generally, the iron sulphide ores tested responded extremely favourably to increasing oxidation temperatures. Ores with higher sulphur grades such as 167 and 201 improved dramatically in terms of sulphur released into solution, whereas ore 169 only showed slight improvement which may not have been statistically significant. Ore 165 showed a slight decrease in gold extraction with increased oxidation temperature. The gold extractions observed at elevated temperatures for these ores saw improvements over tests conducted at baseline temperatures – 91.8% at 55°C compared to an average value of 89.6% at 35°C for sample 167, 73.1% at 55°C compared to 72.3% at 35°C for sample 169, and 90.7% at 55°C compared to 84.0% at 35°C for sample 201. Cyanide consumption recorded for all tests yielded results with a high decree of scatter. Generally, cyanide consumption decreased as caustic consumption increased. Where thiocyanate testing was performed, it was found that as oxidation levels increased, so did the amount of cyanide consumption by production of thiocyanate. As more caustic was added to the leach, cyanate production in the leach slurry drastically decreased, while thiocyanate production increased slightly; these behaviours led to an overall decrease of cyanide usage over the course of oxidation tests. Cyanide consumption varied with regards to rates of air flow during the oxidation phase for each ore. With ore sample 165, the amount of cyanide consumed increased as air input increased, but the opposite effect was observed with the other three tested ores. No clear trend was seen for thiocyanate production with respect to rates of air flow for any ore, due to the different nature of all tested samples. Cyanide consumption increased with oxidation temperatures for all ores. The trends for ores 169 and 201 show that most cyanide consumption is due to cyanide oxidation to cyanate. As temperatures increase, 131  cyanide consumption due to thiocyanate evolution decreases for ore 169 but increases with temperature elevation for ore 201. The oxidation by-products found through XRD analysis were mostly goethite, hematite, and magnetite. Other iron oxides and hydroxides were also observed. For tests that resulted in higher extents of oxidation, more surface products with ferric iron were realised, while tests with little or no sulphide oxidation saw more iron oxides and ferrous iron hydroxides in their tails residues. No gold, silver, or other metals were seen in significant quantities, but unreacted pyrite and marcasite were detected in almost all samples. Economically, the net revenue benefit earned per tonne of ore processed increased an average of $16.49 from baseline cyanidation to baseline oxidation, and higher benefits can be achieved for ores 167 and 201, after high temperature oxidation. All test conditions were found to be economically viable. Operators performing this kind of alkaline oxidation at atmospheric temperatures can apply the results of this work to test their own ore samples. For comparable ores, in most cases increased temperatures and increased rates of airflow have been shown to improve the total extent of oxidation of iron sulphide ores and increase gold extractions compared to samples not treated prior to cyanidation. However, this work focused on a few specific samples from the White Mountain mine, and so tests must be conducted to verify whether iron sulphide ores are similar in composition to these samples, and whether future ores react in the same manner as described in this work. Overall, this research gives a general understanding of the behaviour of such ores with physical parameters commonly used in mineral processing. The research conducted here also verifies that an increase in gold extractions, net gold revenues that can be obtained from each tonne of this type of refractory iron sulphide gold ore with this type of oxidative pre-treatment. Additionally, this work shows that, apart from a few exceptions, all tested physical parameters increase the desired metrics of sulphur oxidation, dissolution of solid ore, and gold extractions in response to conventional cyanidation. This work was originally undertaken with the goals described in Chapter 1. They are given below: 1. To record rates of oxidation of ore samples obtained from the White Mountain mine and relating this back to the varying contents of pyrite, marcasite, arsenopyrite, silicate material, as well as any other significant gold-containing minerals in each ore. Specifically: a. Measure the rates of reaction and develop models to confirm the mode of oxidation b. Determine the controlling mechanism of the reaction and distinguish between factors that control the rate of reaction (e.g. pyrite and marcasite oxidation conforming to various leaching models such as mass-transfer through ion-diffusion, or chemical reaction control) 132  c. Determine the activation energies for iron sulphide minerals in each sample of ore that is tested 2. To delineate the effects of alkaline oxidative pre-treatment on the extraction of gold from each natural ore sample by cyanidation. Parameters investigated include caustic dosage, air sparging levels, and temperature. 3. To describe the oxidation products that resulted from the alkaline pre-treatment step, and any influence they may have on the overall extraction of gold Based on the conclusions given above, the second and third goals described above have been accomplished. The effects of physical parameters on extent of oxidation as well as gold extraction have been elucidated. The oxidation products and surface layers that were found through XRD analysis and their possible effects on gold extraction have been described. The first goal, however, remains uncompleted. To conduct precise surface oxidation and kinetic studies on iron-sulphide minerals such as pyrite, marcasite, and arsenopyrite, pure mineral samples must be used. Each sample must contain only that mineral of interest, without any impurities as far as possible. The rate of oxidation, method/mode of oxidation, controlling mechanism, and activation energy for oxidation of that mineral can then be investigated without any external effects. Thus, there are objectives that can be completed with further testing that will be described below. 5.2 Recommendations for Future Work  Several recommendations can be made for future test work in this area: - Perform a diagnostic leach on all tested ores in this work, as well as those not chosen for oxidation testing to determine the distribution of gold associated with iron sulphide minerals and silica or clay components of each ore - Repeat the tests performed in this work to confirm oxidation and gold extraction results, especially where the nugget effect may have influenced gold extractions, or where the percent differences for results of the same type of test were high (> 5%) - Repeat the tests performed in this work at different alkaline pH values (pH>7) to show how behaviour of the tested ores changes with pH, rather than caustic consumption (which cannot be considered an independent variable using the methodology in this research); after preliminary 133  testing occurs, the pH values which show the best results can be used as “baseline” conditions for future testing - Use pure mineral samples of pyrite, marcasite, and arsenopyrite to investigate the actual oxidation rate, oxidation mode, controlling mechanism, and activation energy of all oxidation reactions that occur under alkaline atmospheric conditions; repeat the process with mineral samples from other parts of the world. Finally, using a modified diagnostic leach method, purify/concentrate the iron-sulphide mineral from the White Mountain ores for kinetic studies - Investigate the effects of silver content on alkaline oxidation of iron sulphides, since a galvanic effect may catalyze oxidation of pyrite, marcasite, and arsenopyrite, despite a small assayed silver head grade being found in each ore (Dreisinger, 2014) - Use Ion Chromatography (IC) or UV-Vis Spectroscopy testing on aqueous samples throughout alkaline oxidation tests to determine the aqueous sulphur species present during oxidation, as well as the possible cyanide complexes or products (such as cyanate or thiocyanate) during the gold leaching phase; increasing the number of samples taken throughout the first two hours of the test would give a better understanding of the incipient oxidation that occurs - Expand the experimental range of variables tested in this work (caustic dosage, air sparging rates, oxidation temperature) so that a clearer understanding could be made of the trends observed with respect to the physical parameters such as air sparging or oxidation temperature; also, expanding the number of physical parameters tested to include things like initial solids %, stirring rate, or P80 grind size could determine which physical parameters influence oxidation of iron sulphides most - In addition to XRD analysis, optical- and Scanning Electron Microscopy could be used on the untreated samples as well as the leached samples of ores, which would further determine not only the characteristic oxidation by-products present as surface layers on mineral particles, but also their porosity, morphology, and thickness. This would provide more concrete information about the effect of iron-oxide/hydroxide/oxy-hydroxide precipitation on gold extractions - Determine a better way of performing these kinds of leaching tests that will not require removal of solids from the reactor at the end of each phase and a better method of running leach tests that will not introduce any water to the reactor via washing of data-recording probes - Perform a solid/liquid separation phase at the end of oxidation tests to discover whether removal of the oxidation leachate will enhance gold leaching results, or mitigate cyanide loss through production of thiocyanate 134  References  Afenya, P. M. (1991). Treatment of Carbonaceous Refractory Gold Ores. Minerals Engineering, 4(7-11), 1043-1055. Ahlberg, E., Forssberg, K. S., & Wang, X. (1990). The Surface Oxidation of Pyrite in Alkaline Solution. Journal of Applied Electrochemistry, 20, 1033-1039. ALS Ammtec Sydney. (2012, April). INVESTIGATION INTO FLOTATION AND CAUSTIC PRETREATMENT OF WHITE MOUNTAIN ORE. Sydney: Ammtec Unit Trust ABN. Amoah-Forson, B. (1986, December). Autoclave Oxidation of Arsenical Gold and Silver Ores and Concentrates: Dissolution Behaviour and Kinetics of Arsenopyrite in Alkaline Hydrothermal Systems. Idaho, U. S. A.: University of Idaho. Asta, M. P., Cama, J., & Acero, P. (2010). Dissolution Kinetics of Marcasite at Acidic pH. European Journal of Mineralogy, 22, 49-61. Asta, M. P., Cama, J., Ayora, C., Acero, P., & de Giudici, G. (2010). Arsenopyrite Dissolution Rates in O2-Bearing Solutions. Chemical Geology, 273, 272-285. Asta, M. P., Perez-Lopez, R., Roman-Ross, G., Illera, V., Cama, J., Cotte, M., & Tucoulou, R. (2013). Analysis of the Iron Coatings Formed During Marcasite and Arsenopyrite Oxidation at Neutral-Alkaline Conditions. Geologica Acta, 11(4), 465-481. Aylmore, M. G. (2005). Alternative Lixiviants to Cyanide for Leaching Gold Ores. In Various, & M. D. Adams (Ed.), Developments in Mineral Processing (Vol. 15: Advances in Gold Ore Processing, pp. 501-539). Elsevier B. V. Aylmore, M. G., & Muir, D. M. (2001). Thiosulfate Leaching of Gold - A Review. Minerals Engineering, 14(2), 135-174. Barrick Gold Corporation. (2014). Goldstrike. Retrieved from Barrick Gold: http://www.barrick.com/operations/united-states/goldstrike/ Barrick Gold Corporation. (2017, February 22). Operations and Technical Update 2017. Retrieved 2017, from Barrick Gold: http://barrick.q4cdn.com/808035602/files/presentation/2017/Barrick-2017-Operations-and-Technical-Update.pdf#page=42 135  Beattie, M. J., & Poling, G. W. (1987). A Study of the Surface Oxidation of Arsenopyrite Using Cyclic Voltammetry. International Journal of Mineral Processing, 20, 87-108. Bhakta, P., Langhans, J. W., & Lei, K. P. (1989). Alkaline Oxidative Leaching of Gold-Bearing Arsenopyrite Ores. U.S. Department of the Interior, Bureau of Mines, Washington, D. C. Bonnissel-Gissinger, P., Alnot, M., Ehrhardt, J.-J., & Behra, P. (1998). Surface Oxidation of Pyrite as a Function of pH. Environmental Science & Technology, 32(19), 2839-2845. Brown, A. D., & Jurinak, J. J. (1989). Mechanism of Pyrite Oxidation in Aqueous Mixtures. Journal of Environmental Quality, 18, 545-550. Buckley, A. N., & Walker, G. W. (1988). The Surface Composition of Arsenopyrite Exposed to Oxidizing Environments. Applied Surface Science, 35, 227-240. Caldeira, C. L., Ciminelli, V. S., Dias, A., & Osseo-Asare, K. (2003). Pyrite Oxidation in Alkaline Solutions: Nature of the Product Layer. International Journal of Mineral Processing, 72, 373-386. Chandra, A. P., & Gerson, A. R. (2010). The Mechanisms of Pyrite Oxidation and Leaching: A Fundamental Perspective. Surface Science Reports, 65, 293-315. Ciminelli, V. S. (1987, May). Oxidation of Pyrite in Alkaline Solutions and Heterogeneous Equilibria of Sulfur- and Arsenic-containing Minerals in Cyanide Solutions. Pennsylvania, United States of America: The Pennsylvania State University. Ciminelli, V. S., & Osseo-Akare, K. (1995, April). Kinetics of Pyrite Oxidation in Sodium Carbonate Solutions. Metallurgical and Maerials Transactions B, 26B, 209-218. Ciminelli, V. S., & Osseo-Akare, K. (1995, August). Kinetics of Pyrite Oxidation in Sodium Hydroxide Solutions. Metallurgical and Materials Transactions, 26B, 677-685. Corkhill, C. L., & Vaughan, D. J. (2009). Arsenopyrite Oxidation - A Review. Applied Geochemistry, 24, 2342-2361. Creamer Media. (2012, April 13). White Mountain Mine, China. Retrieved from Mining Weekly Magazine: http://www.miningweekly.com/article/white-mountain-mine-china-2012-02-17 Crystal Impact. (2018). Match! Phase Identification from Powder Diffraction. (3.6.0.111). Bonn, Germany. Crystallography365. (2014). Gold! The Crystal Structure of Success. Retrieved from Crystallography365: https://crystallography365.wordpress.com/2014/01/17/gold-the-crystal-structure-of-success/ 136  Dasgupta, R., Guan, Y. C., & Han, K. N. (1997, February). The Electrochemical Behaviour of Gold in Ammoniacal Solutions at 75C. Metallurgical and Materials Transactions B. Process Metallurgy and Materials Processing, 28B(1), 5-12. Demopoulos, G. P. (2009). Aqueous Precipitation and Crystallization for the Production of Particulate Solids with Desired Properties. Hydrometallurgy, 96, 199-214. Deschenes, G. (2005). Advances in the Cyanidation of Gold. In Various, & M. D. Adams (Ed.), Developments in Mineral Processing (Vol. 15: Advances in Gold Ore Processing, pp. 479-500). Elsevier B. V. Dreisinger, D. B. (2014). Personal Communication. Eldorado Gold Corporation. (2016). Operations - White Mountain. Retrieved from Eldorado Gold Corp. Website: http://www.eldoradogold.com/assets/operations-and-projects/asia/operations/white-mountain/default.aspx Espiell, F., Roca, A., Cruells, M., & Nunez, C. (1986). Gold and SIlver Recovery by Cyanidation of Arsenopyrite Ore. Hydrometallurgy, 16, 141-151. Fan, C. W., Markuszewski, R., & Wheelock, T. D. (1984). Behaviour of Mineral Matter During Alkaline Leaching of Coal. ACS Division of Fuel Chemistry, 29(4), 319-325. Farley, L. M. (1998, August). A Survey of Conventional and Novel Processes for the Treatment of Refractory Gold. Vancouver, British Columbia, Canada: The University of British Columbia (UBC). Fernandez, P. G., Linge, H. G., & Wadsley, M. W. (1996). Oxidation of Arsenopyrite (FeAsS) in Acid Part I: Reactivity of Arsenopyrite. Journal of Applied Electrochemistry, 26, 575-583. Fernandez, P. G., Linge, H. G., & Willing, M. J. (1996). Oxidation of Arsenopyrite (FeAsS) in Acid Part II: Stoichiometry and Reaction Scheme. Journal of Applied Electrochemistry, 26, 585-591. Filippou, D., & Demopoulos, G. P. (1997, December). Arsenic Immobilization by Controlled Scorodite Precipitation. Journal of the Minerals, Metals, and Materials Society (JOM), 52-55. Fleming, C. A. (2010, May). Basic Iron Sulfate - A Potential Killer in the Processing of Refractory Gold Concentrates by Pressure Oxidation. Minerals and Metallurgical Processing, 27(2), 81-88. Fleming, C. A., McMullen, J., Thomas, K. G., & Wells, J. A. (2003, February). Recent Advances in the Development of an Alternative to the Cyanidation Process: Thiosulfate Leaching and Resin in Pulp. Minerals and Metallurgical Processing, 20(1), 1-9. 137  Fraser, K. S., Walton, R. H., & Wells, J. A. (1991). Processing of Refractory Gold ores. Minerals Engineering, 4(7-11), 1029-1041. Fujita, T., Fujieda, S., Shinoda, K., & Suzuki, S. (2012). Environmental Leaching Characteristics of Scorodite Synthesized with Fe(II) Ions. Hydrometallurgy, 111-112, 87-102. Fujita, T., Taguchi, R., Abumiya, M., Matsumoto, M., Shibata, E., & Nakamura, T. (2008). Novel Atmospheric Scorodite Synthesis by Oxidation of Ferrous Sulfate Solution Part 1. Hydrometallurgy, 90, 92-102. Fujita, T., Taguchi, R., Abumiya, M., Matsumoto, M., Shibata, E., & Nakamura, T. (2009). Effect of pH on Scorodite Synthesis by Oxidation of Ferrous Ions: Physical Properties and Stability of the Scorodite. Hydrometallurgy, 96, 189-198. Gold Price Group. (2017). Gold Price Chart. Retrieved 2017, from Goldprice.org: https://goldprice.org/gold-price-chart.html Goldhaber, M. B. (1983, March). Experimental Study of Metastable Sulfur Oxyanion Formation During Pyrite Oxidation at pH 6-9 and 30C. American Journal of Science, 283, 193-217. Guan, Y. C., & Han, K. N. (1996, June). The Electrochemical Study on the Dissolution Behaviour of Gold in Ammoniacal Solutions at Temperatures Above 100C. Journal of the Electrochemical Society, 143(6), 1875-1880. Guzman, L., Segarra, M., Chimenos, J. M., Cabot, P. L., & Espiell, F. (1999). Electrochemistry of Conventional Gold Cyanidation. Electrochemica Acta, 44, 2625-2632. International Cyanide Management Institute. (2012). Environmental & Health Effects. Retrieved 2014, from International Cyanide Management Code for the Gold Mining Industry: http://www.cyanidecode.org/cyanide-facts/environmental-health-effects International Union of Pure and Applied Chemistry. (2014). Compendium of Chemical Terminology. International Union of Pure and Applied Chemistry. Retrieved from http://goldbook.iupac.org/pdf/goldbook.pdf Jara, J. O., & Bustos, A. A. (1992). Effect of Oxygen on Gold Cyanidation: Laboratory Results. Hydrometallurgy, 30, 195-210. 138  Jones, L. (1998, August). The Biooxidation and Cyanidation of a Refractory Arsenical Gold Concentrate: Factors Influencing Cyanide Consumption. Vancouver, British Columbia, Canada: The University of British Columbia. Klein, B. (2010, Jan.). Processing of Precious Metal Ores. Vancouver, B.C., Canada: The University of British Columbia. Komnitsas, C., & Pooley, F. D. (1989). Mineralogical Characteristics and Treatment of Refractory Gold Ores. Mineral Engineering, 2(4), 449-457. Kondos, P. D., Deschenes, G., & Morrison, R. M. (1995). Process Optimization Studies in Gold Cyanidation. Hydrometallurgy, 39, 235-250. Koslides, T., & Ciminelli, V. S. (1992). Pressure Oxidation of Arsenopyrite and Pyrite in Alkaline Solutions. Hydrometallurgy, 30, 87-106. La Brooy, S. R., Linge, H. G., & Walker, G. S. (1994). Review of Gold Extraction from Ores. Minerals Engineering, 7(10), 1213-1241. Laitos, J. G. (2012, February). The Current Status of Cyanide Regulations. Engineering and Mining Journal, 213(2), 34-40. Langmuir, D., Mahoney, J., & Rowson, J. (2006). Solubility Products of Amorphous Ferric Arsenate and Crystalline Scorodite (FeASO4*2H2O) and their Application to Arsenic Behaviour in Buried Mine Tailings. Geochimica et Cosmochimica Acta, 70, 2942-2956. Lengke, M. F., Sanpawanitchakit, C., & Tempel, R. N. (2009). The Oxidation and Dissolution of Arsenic-Bearing Sulfides. The Canadian Mineralogist, 47(3), 593-613. Logsdon, M. J., Hagelstein, K., & Mudder, T. I. (1999). The Management of Cyanide in Gold Extraction. Ottawa: International Council on Metals and the Environment. Lowson, R. T. (1982, October). Aqueous Oxidation of Pyrite by Molecular Oxygen. Chemical Reviews, 82(5), 461-497. MacArthur, J. S., Forrest, R. W., & Forrest, W. (1887). U.K. Patent No. 14174. Retrieved from http://archive.org/stream/cyanideprocessit00scherich/cyanideprocessit00scherich_djvu.txt MacArthur, J. S., Forrest, R. W., & Forrest, W. (1889). U.S.A. Patent No. 403202. Retrieved from http://archive.org/stream/cyanideprocessit00scherich/cyanideprocessit00scherich_djvu.txt 139  Marsden, J. O., & House, C. I. (2006). The Chemistry of Gold Extraction (2nd ed.). Littleton, Colorado, U.S.A.: Society of Mining, Metallurgy, and Exploration, Inc. Miller, P., & Brown, A. (2005). Bacterial Oxidation of Refractory Gold Concentrates. In Various, & M. D. Adams (Ed.), Developments in Mineral Processing (Vol. 15: Advances in Gold Ore Processing, pp. 371-402). Elsevier B. V. Mineralogical Society of America. (2003). Handbook of Mineralogy (2nd ed., Vol. 1). (J. W. Anthony, R. A. Bideaux, K. W. Bladh, & M. C. Nichols, Eds.) Chantilly, Virginia, U.S.A.: Mineralogical Society of America. Retrieved from http://www.handbookofmineralogy.org/ Mishra, K. K., & Osseo-Asare, K. (1988). Aspects of the Interfacial Electrochemistry of Semiconductor Pyrite (FeS2). Journal of the Electrochemical Society, 135(10), 2502-2509. Moses, C. O., & Herman, J. S. (1991). Pyrite Oxidation at Circumneutral pH. Geochimica et Cosmochimica, 55, 471-482. Muir, D. M., & Aylmore, M. G. (2005). Thiosulfate as an Alternative Lixiviant to Cyanide for Gold Ores. In Various, & M. D. Adams (Ed.), Developments in Mineral Processing (Vol. 15: Advances in Gold Ore Processing, pp. 541-560). Elesevier B. V. Multi Mix Systems Pty. Ltd. (2009). Treatment of Ores Containing Reactive Iron Sulphides. Shelley. Nesbitt, H. W., Muir, I. J., & Pratt, A. R. (1995). Oxidation of Arsenopyrite by Air and Air-saturated, Distilled Water, and Implications for Mechanism of Oxidation. Geochimica et Cosmochimica Acta, 59(9), 1773-1786. Nicholson, R. V., Gillham, R. W., & Reardon, E. J. (1988). Pyrite Oxidation in Carbonate-Buffered Solution: 1. Experimental Kinetics. Geologica et Cosmochimica Acta, 52, 1077-1085. Nicholson, R. V., Gillham, R. W., & Reardon, E. J. (1990). Pyrite Oxidation in Carbonate-Buffered Solution: 2. Rate Control by Oxide Coatings. Geochimica et Cosmochimica, 54, 395-402. Nicol, M. J. (1980). The Anodic Behaviour of Gold Part II - Oxidation in Alkaline Solutions. Gold Bulletin, 13(3), 105-111. Nicol, M. J., & Guresin, N. (2003). Anodic Behaviour of Arsenopyrite and Cathodic Reduction of Ferrate (VI) and Oxygen in Alkaline Solutions. Journal of Applied Electrochemistry, 33, 1017-1024. Nicol, M. J., Miki, H., & Basson, P. (2013). The Effects of Sulphate Ions and Temperature on the Leaching of Pyrite 2. Dissolution Rates. Hydrometallurgy, 133, 182-187. 140  Nicol, M. J., Miki, H., Zhang, S., & Basson, P. (2013). The Effects of Sulphate Ions and Temperature on the Leaching of Pyrite 1. Electrochemistry. Hydrometallurgy, 133, 188-196. Outokompu Research Oy. (2006). HSC Chemistry 6.0 Software. Pori, Finland. Paktunc, D., Dutrizac, J., & Gertsman, V. (2008). Synthesis and Phase Transformations Involving Scorodite, Ferric Arsenate and Arsenical Ferrihydrite: Implications for Arsenic Mobility. Geochimica et Cosmochimica Acta, 72, 2649-2672. Papangelakis, V. G. (1986, July). Aqueous Pressure Oxidation of Arsenopyrite. Montreal, Quebec, Canada: McGill University. Papangelakis, V. G., & Demopoulos, G. P. (1990). Acid Pressure Oxidation of Arsenopyrite: Part I, Reaction Chemistry. Canadian Metallurgical Quarterly, 29(1), 1-12. Papangelakis, V. G., & Demopoulos, G. P. (1990). Acid Pressure Oxidation of Arsenopyrite: Part II, Reaction Kinetics. Canadian Metallurgical Quarterly, 29(1), 13-20. Papangelakis, V. G., & Demopoulos, G. P. (1991). Acid Pressure Oxidation of Pyrite: Reaction Kinetics. Hydrometallurgy, 26, 309-325. Prasad, M. S., Mensah-Biney, R., & Pizarro, R. S. (1991). Modern Trends in Gold Processing - Overview. Minerals Engineering, 4(12), 1257-1277. Puigdomenech, I. (2010). Make Equilibrium Diagrams Using Sophisticated Algorithms (MEDUSA) Software. MEDUSA. Stockholm, Sweden: KTH Royal Institute of Technology. Qi, P. H., & Hiskey, J. B. (1991). Dissolution Kinetics of Gold in Iodide Solutions. Hydrometallurgy, 27, 47-62. Ruitenberg, R., Hansford, G. S., Reuter, M. A., & Breed, A. W. (1999). The Ferric Leaching Kinetics of Arsenopyrite. Hydrometallurgy, 52, 37-53. Sand, W., Gehrke, T., Jozsa, P.-G., & Schippers, A. (2001). (Bio)chemistry of Bacterial Leaching - Direct vs. Indirect Bioleaching. Hydrometallurgy, 59, 159-175. Schippers, A., & Sand, W. (1999, January). Bacterial Leaching of Metal Sulfides Proceeeds by Two Indirect Mechanisms via Thiosulfate or via Polysulfides and Sulfur. Applied and Environmental Microbiology, 65(1), 319-321. Schippers, A., Jozsa, P.-G., & Sand, W. (1996, September). Sulfur Chemistry in Bacterial Leaching of Pyrite. Applied and Environmental Microbiology, 62(9), 3424-3431. 141  Schippers, A., Rohwerder, T., & Sand, W. (1999). Intermediary Sulfur Compounds in Pyrite Oxidation: Implications for Bioleaching and Biodepyritization of Coal. Applied Microbiology Biotechnology, 52, 104-110. Senanayake, G. (2004). Gold Leaching in Non-Cyanide Lixiviant Systems: Critical Issues on Fundamentals and Applications. Minerals Engineering, 17, 785-801. Senanayake, G. (2005). Kinetics and Reaction Mechanism of Gold Cyanidation: Surface Reaction Model via Au(I)-OH-CN Complexes. Hydrometallurgy, 80, 1-12. Smyth, J. R., & McCormick, T. C. (1994). Crystallographic Data for Minerals. In T. J. Ahrens, Mineral Physics and Crystallography: A Handbook of Physical Constants. U.S.A.: American Geophysical Union. Sun, H., Chen, M., Zou, L., Shu, R., & Ruan, R. (2015). Study of the Kinetics of Pyrite Oxidation Under Controlled Redox Potential. Hydrometallurgy, 155, 13-19. Swaminathan, C., Pyke, P., & Johnston, R. F. (1993). Reagent Trends in the Gold Extraction Industry. Minerals Engineering, 6(1), 1-16. Tao, D. P., Li, Y. Q., Richardson, P. E., & Yoon, R. H. (1994). The Incipient Oxidation of Pyrite. Colloids and Surfaces A: Physiochemical and Engineering Aspects, 93, 229-239. The Chemical Rubber Company. (2014). CRC Handbook of Chemistry and Physics (95th ed.). (W. M. Haynes, Ed.) Boca Raton, Florida, U.S.A.: Taylor and Francis Group. Thomas, K. G. (2005). Pressure Oxidation Review. In Various, & M. D. Adams (Ed.), Developments in Mineral Processing (Vol. 15: Advances in Gold Ore Processing, pp. 345-369). Elsevier B. V. Tossel, A., Vaughan, D. J., & Burdett, J. K. (1981). Pyrite, Marcasite, and Arsenopyrite Type Minerals: Crystal Impact and Structural Principles. Physics and Chemistry of Minerals, 7, 177-184. Uhlig, I., Szargan, R., Nesbitt, H. W., & Laajelehto, K. (2001). Surface States and Reactivity of Pyrite and Marcasite. Applied Surface Science, 179, 222-229. Wadsworth, M. E. (2000). Surface Processes in Silver and Gold Cyanidation. International Journal of Mineral Processing, 58, 351-368. Wadsworth, M. E., Zhu, X., Thompson, J. S., & Pereira, C. J. (2000). Gold Dissolution and Activation in Cyanide Solution: Kinetics and Mechanism. Hydrometallurgy, 57, 1-11. Walker, F. P., Schreiber, M. E., & Rimstidt, J. D. (2006). Kinetics of Arsenopyrite Oxidatitive Dissolution by Oxygen. Geochimica et Cosmochimica Acta, 70, 1668-1676. 142  Williamson, M. A., & Rimstidt, J. D. (1994). The Kinetics and Electrochemical Rate-Determining Step of Aqueous Pyrite Oxidation. Geochimica et Cosmochimica Acta, 58(24), 5443-5454. Yannopoulos, J. C. (1991). The Extractive Metallurgy of Gold (1st ed.). New York, New York, U.S.A.: Van Nostrand Reinhold. Young, C. A., & Jordan, T. S. (1995). Cyanide Remediation: Current and Past Technologies. Proceedings of the 10th Annual Conference on Hazardous Waste Research (pp. 104-129). Manhattan, Kansas: Department of Metallurgical Engineering, Montana Tech. Yu, Y., Zhu, Y., Gao, Z., Gammons, C. H., & Li, D. (2007). Rates of Arsenopyrite Oxidation by Oxygen and Fe(III) at pH 1.8 - 12.6 and 15 - 45C. Environmental Science and Technology, 41(18), 6460-6464. Zhang, S. (2004, March). Oxidation of Refractory Gold Concentrates and Simultaneous Dissolution of Gold in Aerated Alkaline Solutions. Western Australia, Australia: Murdoch University. Zhang, S., & Nicol, M. J. (2003). An Electrochemical Study of the Dissolution of Gold in Thiosulfate Solutions Part 1: Alkaline Solutions. Journal of Applied Electrochemistry, 33, 767-775. Zhang, S., & Nicol, M. J. (2005). An Electrochemical Study of the Dissolution of Gold in Thiosulfate Solutions Part II: Effect of Copper. Journal of Applied Electrochemistry, 35, 339-345. Zhang, Y., Fang, Z., & Muhammed, M. (1997). On the Solution Chemistry of Cyanidation of Gold and Silver Bearing Sulphide Ores. A Critical Evaluation of Thermodynamic Calculations. Hydrometallurgy, 46, 251-269. Zhu, X., Li, J., & Wadsworth, M. E. (1992). Kinetics of the Transpassive Oxidation of Pyrite. Utah University, Department of Metallurgical Engineering. Washington D. C.: U.S. Department of Energy. Retrieved from https://www.osti.gov/scitech/servlets/purl/10128685-xlcB5i/ Zhu, X., Li, J., & Wadsworth, M. E. (1994). Characterization of Surface Layers Formed During Pyrite Oxidation. Colloids and Surfaces A: Physiochemical and Engineering Aspects, 93, 201-210.   143  Appendices  Appendix A: Analytical Methods Used to Test Leaching Samples  Brief descriptions of the analytical tests performed on aqueous and solid samples taken during each leaching test are given below. A.1 Gold Fire Assays  This test was performed on “head” samples (untreated and un-leached ground ore samples) as well as on every “tails” sample (leached residue, filtered and dried after cyanidation). These tests were performed by the metallurgical laboratory at Inspectorate International Ltd., in Richmond, B.C. The test involved burning the sample with an extremely high temperature flame, in the presence of various other compounds. These compounds act to produce a slag, removing impurities from the burned sample, leaving only heavier elements such as lead, gold, silver, platinum, and palladium in the form of a metallic bead at the bottom of the crucible. The bead is then further heated in a porous container called a cupel, which is typically ceramic. The lead bead separates due to oxidation of the lead, leaving only precious metals inside the container. Finally, the precious metal bead is dissolved, and chemical analysis is performed. Typically, ICP-AES/OES analysis is used to detect the presence and concentration of gold, however other techniques may also be used, depending on which precious metals are being analyzed. A.2 Sulphate Tests  This test was performed by AuTec Innovative Extractive Solutions Ltd. in Vancouver, B.C. ICP-AES/OES was performed after acidifying samples with hydrochloric acid (HCl) and the sulphur content in the sample was totalled and divided by the mass ratios of sulphate (SO4-2) and sulphur (S), which is approximately 3.00. While these tests were performed to analyze for sulphate, their results may be more accurately interpreted as aqueous sulphur, using the mass ratios to multiply results by three to get total sulphur in solution. Additionally, other elements were analyzed, such as iron, calcium, aluminum, and magnesium, as part of the ICP-AES/OES analysis. A.3 Solid Sulphur Tests  These tests were performed using LECO instruments, which burned solid residues taken from the 8-hour and 24-hour points in the leaching test (at the end of the 8-hour oxidation phase) and analyzing the off-gas that was volatilized. The detection method was Infrared Analysis (IR), and the instrument was then able to 144  determine the total amount of sulphur that was present in the sample. A potential downside to this method can arise if sulphur is somehow not burned in the reaction chamber, due to occlusion of silicates or other gangue (foreign) matter. However, the potential for this possibility is quite low, as the sample residues came from previous oxidized material that was already finely ground (~50 μm). These tests were performed by Inspectorate in Richmond, B. C. A.4 Aqueous Gold Tests  These tests were performed by Inspectorate International Ltd. in Richmond, B.C. The tests were performed using a technique called Atomic Adsorption Spectroscopy, whereby aqueous sample is burned in a temperature-controlled flame, and the volatilized off-gas is analyzed using a detector called a spectrometer. In this technique, a lamp emits light or radiation which is adsorbed by the atoms of the gaseous sample; the un-adsorbed light passes through a monochromator, which reduces the signal noise, down to one wavelength. Finally, an optical detector, called a spectrometer, captures the light and the instrument then matches the signal received to various elements in its database. This technique is one of the standard method by which aqueous gold is analyzed in samples. Preparations for analysis involve ensuring a clean burning apparatus, and a well-adjusted and temperature-controlled flame, appropriate dilution of the solution sample, and calibration of the spectrometer machine to capture the concentration of the element accurately. A.5 Thiocyanate Tests  These tests were performed by Inspectorate International Ltd. in Richmond, B.C. using a Spectrophotometer Colourimetry analysis technique. This technique uses the Beer-Lambert law to obtain a colour spectrum which can then be used to determine individual compounds and their concentrations. A similar technique – UV-Vis Spectroscopy develops a spectrum by passing UV light through a sample; the SC technique uses visible light. Some advantages of using the SC technique are that samples can be analyzed without prior dilution, and the concentrations of multiple compounds can be accurately found. Thiocyanate levels are not typically sensitive to ageing; the compound is generally stable and does not normally decompose. However, there is potential to artificially increase the concentration of thiocyanates in solution in the presence of sulphur or partially oxidized sulphur compounds such as thiosulphate, sulphite, or sulphate. Care must be taken to avoid contamination and obtain clear solution samples for analysis. For this work, solutions analyzed were taken from the PLS at the end of cyanidation tests. All solutions were clear and those at the end of tests which included pre-oxidation had sulphur that was fully oxidized to sulphate, as evidenced by typically high sulphate levels at the end of the pre-treatment phase.145  Appendix B: XRD Imaging Results for Untreated Samples B.1 Sample 165  Figure 67: XRD Pattern for untreated ("Head") sample 165; analysis and Rietveld Refinement performed by the department of Earth and Ocean Sciences at UBC 2Th Degrees8075706560555045403530252015105Sqrt(Counts)220200180160140120100806040200-201AR_165.raw_1 Quartz low 90.86 %Barite 6.18 %Kaolinite 1A 0.80 %Iron-alpha 0.40 %Magnetite 0.12 %Illite/Muscovite 2M1 1.28 %Hematite 0.35 %146  B.2 Sample 167  Figure 68: XRD pattern for untreated ("Head") sample 167; analysis and Rietveld Refinement performed by the department of Earth and Ocean Sciences at UBC 2Th Degrees8075706560555045403530252015105Sqrt(Counts)220200180160140120100806040200-202AR_167.raw_1 Quartz low 78.00 %Pyrite 5.57 %Marcasite 4.48 %Gypsum 2.36 %Melanterite 2.52 %Kaolinite 1A 1.78 %Barite 1.83 %Illite/Muscovite 2M1 2.72 %Goyazite 0.74 %147  B.3 Sample 169  Figure 69: XRD pattern for untreated ("Head") sample 169; analysis and Rietveld Refinement performed by the department of Earth and Ocean Sciences at UBC 2Th Degrees8075706560555045403530252015105Sqrt(Counts)200180160140120100806040200-203AR_169.raw_1 Quartz low 65.33 %Pyrite 11.04 %Marcasite 12.74 %Gypsum 3.25 %Illite/Muscovite 2M1 4.27 %Kaolinite 1A 1.48 %Sulfur 1.06 %Goyazite 0.31 %Jarosite 0.53 %148  B.4 Sample 201  Figure 70: XRD pattern for untreated ("Head") sample 201; analysis and Rietveld Refinement performed by the department of Earth and Ocean Sciences at UBC 2Th Degrees8075706560555045403530252015105Sqrt(Counts)220200180160140120100806040200-204AR_201.raw_1 Quartz low 85.70 %Pyrite 3.07 %Gypsum 2.79 %Illite/Muscovite 2M1 2.88 %Kaolinite 1A 1.78 %Jarosite 0.75 %Actinolite 0.33 %Marcasite 0.75 %Goyazite 0.27 %Barite 1.68 %149  Appendix C: Sample Calculations  Given below are explanations and sample calculations for various metrics determined in this work. C.1 Gold Extraction  Gold extraction is calculated via the following formula: % 𝑇𝑜𝑡𝑎𝑙 𝐸𝑥𝑡𝑟𝑎𝑐𝑡𝑖𝑜𝑛 𝑜𝑓 𝐺𝑜𝑙𝑑 =𝑇𝑜𝑡𝑎𝑙 𝑔𝑜𝑙𝑑 𝑖𝑛 𝑠𝑜𝑙𝑢𝑡𝑖𝑜𝑛𝑇𝑜𝑡𝑎𝑙 𝑔𝑜𝑙𝑑 𝑖𝑛 𝑡𝑒𝑠𝑡𝑒𝑑 𝑠𝑎𝑚𝑝𝑙𝑒𝑥100    Eq. 106a % 𝑇𝑜𝑡𝑎𝑙 𝐸𝑥𝑡𝑟𝑎𝑐𝑡𝑖𝑜𝑛 𝑜𝑓 𝐺𝑜𝑙𝑑 = 1 −𝑇𝑜𝑡𝑎𝑙 𝑔𝑜𝑙𝑑 𝑖𝑛 𝑡𝑎𝑖𝑙𝑠 𝑟𝑒𝑠𝑖𝑑𝑢𝑒𝑇𝑜𝑡𝑎𝑙 𝑔𝑜𝑙𝑑 𝑖𝑛 𝑡𝑒𝑠𝑡𝑒𝑑 𝑠𝑎𝑚𝑝𝑙𝑒𝑥100%   Eq. 106b Where the total gold in solution is given by the following equation: 𝐴𝑢𝑠𝑜𝑙𝑢𝑡𝑖𝑜𝑛,𝑡𝑜𝑡𝑎𝑙 = 𝐴𝑢𝑖𝑛𝑡𝑒𝑟𝑚𝑒𝑑𝑖𝑎𝑡𝑒 𝑠𝑎𝑚𝑝𝑙𝑒𝑠 + 𝐴𝑢𝑃𝑟𝑒𝑔𝑛𝑎𝑛𝑡 𝐿𝑒𝑎𝑐ℎ 𝑆𝑜𝑙𝑢𝑡𝑖𝑜𝑛 + 𝐴𝑢𝑊𝑎𝑠ℎ   Eq. 107 And the total gold present in a tested sample is given by the following equation: 𝐴𝑢𝑠𝑎𝑚𝑝𝑙𝑒,𝑡𝑜𝑡𝑎𝑙 = 𝐴𝑢𝑠𝑜𝑙𝑢𝑡𝑖𝑜𝑛,𝑡𝑜𝑡𝑎𝑙 + 𝐴𝑢𝑟𝑒𝑠𝑖𝑑𝑢𝑒       Eq. 108 Therefore, equations 106a and 106b above become: % 𝑇𝑜𝑡𝑎𝑙 𝐸𝑥𝑡𝑟𝑎𝑐𝑡𝑖𝑜𝑛 𝑜𝑓 𝐺𝑜𝑙𝑑 =𝐴𝑢𝑠𝑜𝑙𝑢𝑡𝑖𝑜𝑛,𝑡𝑜𝑡𝑎𝑙𝐴𝑢𝑠𝑎𝑚𝑝𝑙𝑒,𝑡𝑜𝑡𝑎𝑙𝑥100% =𝐴𝑢𝑠𝑜𝑙𝑢𝑡𝑖𝑜𝑛,𝑡𝑜𝑡𝑎𝑙𝐴𝑢𝑠𝑜𝑙𝑢𝑡𝑖𝑜𝑛,𝑡𝑜𝑡𝑎𝑙+𝐴𝑢𝑟𝑒𝑠𝑖𝑑𝑢𝑒𝑥100% Eq. 106c % 𝑇𝑜𝑡𝑎𝑙 𝐸𝑥𝑡𝑟𝑎𝑐𝑡𝑖𝑜𝑛 𝑜𝑓 𝐺𝑜𝑙𝑑 = 1 −𝐴𝑢𝑟𝑒𝑠𝑖𝑑𝑢𝑒𝐴𝑢𝑠𝑜𝑙𝑢𝑡𝑖𝑜𝑛,𝑡𝑜𝑡𝑎𝑙+𝐴𝑢𝑟𝑒𝑠𝑖𝑑𝑢𝑒𝑥100%   Eq. 106d Expanding on all the terms, the total extraction of gold for a sample is given by the following equation % 𝑇𝑜𝑡𝑎𝑙 𝐸𝑥𝑡𝑟𝑎𝑐𝑡𝑖𝑜𝑛 𝑜𝑓 𝐺𝑜𝑙𝑑 =𝐴𝑢𝑖𝑛𝑡𝑒𝑟𝑚𝑒𝑑𝑖𝑎𝑡𝑒 𝑠𝑎𝑚𝑝𝑙𝑒𝑠+𝐴𝑢𝑃𝑟𝑒𝑔𝑛𝑎𝑛𝑡 𝐿𝑒𝑎𝑐ℎ 𝑆𝑜𝑙𝑢𝑡𝑖𝑜𝑛+𝐴𝑢𝑊𝑎𝑠ℎ𝐴𝑢𝑖𝑛𝑡𝑒𝑟𝑚𝑒𝑑𝑖𝑎𝑡𝑒 𝑠𝑎𝑚𝑝𝑙𝑒𝑠+𝐴𝑢𝑃𝑟𝑒𝑔𝑛𝑎𝑛𝑡 𝐿𝑒𝑎𝑐ℎ 𝑆𝑜𝑙𝑢𝑡𝑖𝑜𝑛+𝐴𝑢𝑊𝑎𝑠ℎ+𝐴𝑢𝑟𝑒𝑠𝑖𝑑𝑢𝑒𝑥100%            Eq. 106 The aqueous gold in solution is given by equation 106 above, and the gold present in the tails residue is determined through fire assay. The amount of gold present in intermediate samples, pregnant leach solution, and wash water is determined by the following equation: 𝐴𝑢𝑠𝑎𝑚𝑝𝑙𝑒,𝑖 = 𝐶𝑖𝑉𝑖         Eq. 108 Where Ci is the concentration of gold of a sample taken at time i, determined by Atomic Adsorption Spectroscopy (AAS) analysis, and Vi is the volume of said sample. For the gold found in the pregnant leach solution (PLS) and wash, the concentrations of each of these samples were multiplied by the total volumes 150  of each solution, respectively. To obtain a value for the total gold found in the intermediate samples, a simple sum of all individually calculated gold values is used. An example of these types of calculations are given below: Table 36: Example Gold Calculations Sample or Solution Volume (L) Au Concentration without Dilution Factors (mg/L) Au present (mg) 1 Hour Aliquot 0.00836 1.66 0.01387 2 Hour Aliquot 0.00849 1.56 0.01325 4 Hour Aliquot 0.00853 1.50 0.01280 8 Hour Aliquot 0.00866 1.50 0.01299 PLS 0.535 1.48 0.7918 Wash 0.250 0.10 0.02500 Total   0.8697 Sample Weight (g) Au Concentration (g/mt) Au present (mg) Tails Residue 82.94 2.47 0.2049 Sample Total Gold Present (mg) Solution (intermediate samples + PLS + Wash) 0.8697 Tails Residue 0.2049 Total Gold 1.075  % 𝑇𝑜𝑡𝑎𝑙 𝐸𝑥𝑡𝑟𝑎𝑐𝑡𝑖𝑜𝑛 𝑜𝑓 𝐺𝑜𝑙𝑑 =𝐴𝑢𝑠𝑜𝑙𝑢𝑡𝑖𝑜𝑛,𝑡𝑜𝑡𝑎𝑙𝐴𝑢𝑠𝑎𝑚𝑝𝑙𝑒,𝑡𝑜𝑡𝑎𝑙𝑥100% =0.8697 𝑚𝑔1.075 𝑚𝑔𝑥100% = 80.9%  C.2 Extraction Curves  To obtain extraction curves, it was first necessary to calculate solution dilution factors for each point in time. This is necessary because the leaching solution present in the reactors during each test tended to become more dilute over time (due to the addition of wash water for probes that were used to determine pH, ORP (Eh), Dissolved Oxygen (D.O.), and temperature levels at each sampling time). Since the leaching solution became more dilute with time, the concentrations of aqueous gold in each sample would decrease, even though more aqueous gold was being released into solution as the leaching progressed. 151  Dilution factors compensate for dilution of the leaching solution over time. Dilution factors are calculated by dividing each successive volume of leaching solution by the original volume of leachate to obtain ratios. These ratios are then multiplied with the concentrations of elements or compounds found through assay, ICP, or AAS analysis. Continuing from the previous example, an example is given below: Table 37: Example of Gold Concentrations with Dilution Factors Time (Hours) Solution Volumes (L) Dilution Factors Au Concentration determined by AAS (mg/L) Au Concentration with Dilution Factor (mg/L) 0 0.40767 1.000 0.00 0.00 1 0.46145 1.132 1.66 1.80 2 0.48428 1.188 1.56 1.75 4 0.50697 1.244 1.50 1.73 8 0.52953 1.299 1.50 1.78 24 0.53763 1.319 1.48 1.76  The newly calculated concentrations show only a slight decrease in gold concentration over time, whereas the original values are heavily diluted. From these new numbers, an extraction curve can be developed. Extraction values at any given time are calculated in a similar fashion to the dilution factors – the total extraction calculated above is multiplied by ratios of the concentrations at various points divided by the final dilution-compensated concentration value of gold. The example above continues below: Table 38: Example of Extraction Curve Calculations Time (Hours) Au Concentration with Dilution Factor (mg/L) Extraction Ratios Total Extraction (%) 0 0.00 0.00/1.76 = 0.000 0 x 80.9% = 0.00% 1 1.80 1.80/1.76 = 1.022 1.022 x 80.9% = 80.7% 2 1.75 1.75/1.76 = 0.990 0.990 x 80.9% = 80.2% 4 1.73 1.73/1.76 = 0.981 0.981 x 80.9% = 79.4% 8 1.78 1.78/1.76 = 1.010 1.010 x 80.9% = 81.8% 24 1.76 1.76/1.76 = 1.000 1.000 x 80.9% = 80.9%  This produces the following extraction curve: 152   Figure 71: Example gold extraction curve Similarly, using the newly obtained extraction values for each time point, the aqueous gold in solution at any given time can be calculated by multiplying the Extraction value at that time by the total (summed) aqueous gold in solution. C.3 Calculated Head Values  To calculate the Head value for a given tested sample of ore, simply divide the total gold present in the sample (aqueous gold + gold in tails residue) by the mass of the tested sample, and convert the units to g/mt. Using the numbers above, an example is given below. Mass of tested ore: 84.96g Gold Present in the tested sample (solution + residue): 1.074575 mg 𝐶𝑎𝑙𝑐𝑢𝑙𝑎𝑡𝑒𝑑 𝐻𝑒𝑎𝑑 𝑉𝑎𝑙𝑢𝑒 =(1.075 𝑚𝑔 𝑥 𝑔 𝑔𝑜𝑙𝑑1000 𝑚𝑔)(84.96𝑔 𝑥 𝑡 𝑜𝑟𝑒1000000𝑔)= 12.7 𝑔 𝑔𝑜𝑙𝑑/𝑚𝑡 C.4 Cyanide Consumption  To determine the overall consumption of cyanide for each test, a volumetric titration process was used. A 25.00 mL volume of pregnant leach solution (PLS) was taken after leaching was completed, and then reacted with a 0.051 M silver nitrate (AgNO3) in the presence of a 5% potassium iodide (KI) indicator solution. 0%10%20%30%40%50%60%70%80%90%100%0 5 10 15 20 25Gold Extraction %Time (Hours)Gold Extraction vs. Timew/ DF (%)153  The residual cyanide in the solution reacts quickly with the silver nitrate to produce a silver-cyanide complex. This typically consumes all residual cyanide present in the sample, while excess silver nitrate reacts with the potassium iodide indicator to produce potassium nitrate and silver iodide, after the primary silver leaching reaction. Silver iodide (AgI) is a precipitate with low solubility, with a typically yellow-white, amorphous, and cloudy appearance. Once the solution turns a pale, cloudy yellow, the titration is complete, and the volume of silver nitrate is recorded. The primary and secondary reactions of the titrations are given below: Primary Reaction: 2 NaCN(aq) + AgNO3(aq) = NaAg(CN)2(aq) + NaNO3(aq)   Eq. 109 Secondary Reaction: KI(aq) + AgNO3(aq) = AgI(s) + KNO3(aq)    Eq. 110 Using this information, as well as the volumes of both the silver nitrate solution and the PLS solution, a calculation of the number of moles of cyanide present in the solution can be undertaken, and subsequently conversion into units typically used in industry such as g/L or kg/t. A sample calculation of the consumption of cyanide is given below: Sample weight of ore: 89.74 g Original dosage of sodium cyanide: 9.98 kg/t Solution volume: 25.59 mL 0.051M AgNO3 volume used until titration point: 1.26 mL 0.051 𝑀 𝐴𝑔𝑁𝑂3 𝑥 1.26 𝑚𝐿(1000 𝑚𝐿𝐿)= 0.00006426 𝑚𝑜𝑙𝑒𝑠 𝐴𝑔𝑁𝑂3 𝑐𝑜𝑛𝑠𝑢𝑚𝑒𝑑 𝑏𝑦 𝑟𝑒𝑠𝑖𝑑𝑢𝑎𝑙 𝑐𝑦𝑎𝑛𝑖𝑑𝑒  0.00006426 𝑚𝑜𝑙𝑒𝑠 𝐴𝑔𝑁𝑂3 𝑥 2 𝑚𝑜𝑙𝑒𝑠 𝑁𝑎𝐶𝑁1 𝑚𝑜𝑙𝑒 𝐴𝑔𝑁𝑂3= 0.0001285 𝑚𝑜𝑙𝑒𝑠 𝑁𝑎𝐶𝑁 𝑝𝑟𝑒𝑠𝑒𝑛𝑡  𝑖𝑛 𝑃𝐿𝑆  0.00012852 𝑚𝑜𝑙𝑒𝑠 𝑁𝑎𝐶𝑁 𝑥 49.0072 𝑔1 𝑚𝑜𝑙𝑒 𝑁𝑎𝐶𝑁= 0.006298 𝑔 𝑁𝑎𝐶𝑁  0.006298 𝑔 𝑁𝑎𝐶𝑁(25.59 𝑚𝐿1000 𝑚𝐿/𝐿)= 0.2461 𝑔 𝑁𝑎𝐶𝑁𝐿 𝑠𝑜𝑙𝑢𝑡𝑖𝑜𝑛   0.006298 𝑔 𝑁𝑎𝐶𝑁89.74 𝑔 𝑜𝑟𝑒𝑥 1 𝑘𝑔 𝑁𝑎𝐶𝑁1000 𝑔 𝑁𝑎𝐶𝑁 𝑥 1000000 𝑔 𝑜𝑟𝑒1 𝑡𝑜𝑛𝑛𝑒 𝑜𝑟𝑒= 1.69𝑘𝑔 𝑁𝑎𝐶𝑁𝑡 𝑜𝑟𝑒  For samples that have had ICP analysis for thiocyanates (SCN-), the same sort of calculation is performed, where the concentration of the thiocyanate found in solution is multiplied by the volume of the pregnant leach solution and converted to get the consumption of cyanide by thiocyanates. A sample is shown below: 154  SCN- concentration: 664 mg/L PLS volume: 0.61597 L Thiocyanate Formation Reaction: CN-(aq) + S0(s) = SCN-(aq)     Eq. 111 664𝑚𝑔𝐿𝑥 0.61597 𝐿 𝑃𝐿𝑆 = 409 𝑚𝑔 𝑆𝐶𝑁  409 𝑚𝑔 𝑆𝐶𝑁 𝑥 1 𝑔1000 𝑚𝑔 𝑥 1 𝑚𝑜𝑙𝑒 𝑆𝐶𝑁58.08 𝑔 𝑥 1 𝑚𝑜𝑙𝑒 𝐶𝑁−1 𝑚𝑜𝑙𝑒 𝑆𝐶𝑁 𝑥 1 𝑚𝑜𝑙𝑒 𝑁𝑎𝐶𝑁1 𝑚𝑜𝑙𝑒 𝐶𝑁−= 0.007 𝑚𝑜𝑙𝑒𝑠 𝑁𝑎𝐶𝑁  0.007 𝑚𝑜𝑙𝑒𝑠 𝑁𝑎𝐶𝑁 𝑥  49.0072 𝑔1 𝑚𝑜𝑙𝑒 𝑁𝑎𝐶𝑁= 0.345 𝑔 𝑁𝑎𝑐𝑁 𝑢𝑠𝑒𝑑 𝑓𝑜𝑟 𝑆𝐶𝑁 𝑓𝑜𝑟𝑚𝑎𝑡𝑖𝑜𝑛 0.345 𝑔 𝑁𝑎𝐶𝑁 𝑥 1 𝑘𝑔 𝑁𝑎𝐶𝑁1000 𝑔 𝑁𝑎𝐶𝑁 𝑥 1000000 𝑔 𝑜𝑟𝑒1 𝑡𝑜𝑛𝑛𝑒 𝑜𝑟𝑒= 3.85𝑘𝑔 𝑁𝑎𝐶𝑁 𝑐𝑜𝑛𝑠𝑢𝑚𝑒𝑑𝑡 𝑜𝑟𝑒  Knowing that the original dosage of cyanide for this test was 9.98 kg/t, residual cyanide was calculated as 1.69 kg/t, and 3.85 kg/t of cyanide was consumed to produce thiocyanates, a breakdown of cyanide usage and consumption can be constructed. The consumption of cyanide through gold lixiviation was calculated in the same way as shown above. Typically cyanide usage for precious metal lixiviation is quite small compared to formation of other cyanide compounds such as cyanates, thiocyanates, and other metal-cyanide complexes. C.5 Oxidation  A similar process to that outlined above for gold is used to obtain the total aqueous sulphur in solution and total solid sulphur in the residue at the end of the 8-hour oxidation phase. Oxidation curves once again use dilution factors to compensate for decreasing aqueous sulphur concentrations due to dilution of the leachate during an oxidation trial. As with gold, the total amount of sulphur in the sample (solution + solid), as well as the calculated “Head” values of sulphur are determined with the unchanged concentrations of sulphur in solution, while the extraction curves are calculated using concentrations that have been modified using dilution factors. To obtain the amount of sulphur in solution at any given point, the extraction level at that time is multiplied by the true total amount of sulphur found in the sample. This process gives an extremely close approximation to the actual values found in experiments for both gold and sulphur, and errors between these models and actual values are negligible for most tests, with the bulk of error coming from minor accidental spillage, evaporation, or dilution of less than 10 mL total. 155  Appendix D: Example Mass Balance Table  Given below is an example of the mass balance tables used for all tests performed for this work. For Baseline Cyanidation Tests, the tables showing calculated volumes of leach solution at each sampling time are included, but these tables (for both Oxidation as well as Cyanidation stages) are not always included to save space.  The order of tables for oxidation tests are as follows: 1st page: - List of reagents inputted to the reactor and pH/ORP calibration information - Raw data for oxidation stages, aqueous and solid samples extracted, and assayed sulphur results 2nd page: - Calculation of solution volumes present in the reactor at each sampling time - Results of ICP analysis of aqueous samples and oxidation calculations 3rd page: - Tabulation of caustic sodium hydroxide dosed during each stage - Raw data for cyanidation stage, aqueous samples extracted, filtration results, and tails gold and sulphur assay 4th page: - Calculation of solution volumes present in the reactor at each sampling time - Calculation of gold head grade and gold extraction levels over time  For baseline cyanidation tests, the tables dealing with oxidation are omitted, but the tables involving reagents inputted into the leaching vessel, and sodium hydroxide dosed over time are still shown. The first example given below is the one shown in Appendix C. The second example is an example with oxidation calculations added. 156  D.1 Sample 165 Baseline Cyanidation Test #2   Alkaline Oxidation Ore Types MoistureTest Base Cyan 2 165 0.1096 0.077Date 08-03-2017 167 0.1218 0.1662Ore Type 1 169 0.1392165 201 0.1192Moisture 7.70% 24-04-2017FileAssay Data Phases % Results 17-390-00111Au - Head FA 14.16 g/mt Actinolite 0Ag - ICP-MS 9.50 g/mt Barite 6.2Fe - ICP-MS 1.79 % Goyazite 0Total S (LECO) 0.88 % Gypsum 0Ba - ICP-MS 20188.00 g/mt Hematite 0.3Al - ICP-MS 1.00 % Illite/Muscovite 2M1 1.3As - ICP-MS 172.00 g/mt Iron 0.4Hg - ICP-MS 10.64 g/mt Jarosite 0Kaolinite 0.8Reactor 2 Magnetite 0.1Reactor Weight 1204.8 g Marcasite 0Melanterite 0Needed 85 g Pyrite 0Ore Weighed 92.09 g Quartz 90.9Actual Ore Added 92.05 g Sulfur 0Solids 20% Total 100Dry Ore Added 84.96 g Oxidation ConditionsNaOH Initial Dosage 0 kg/tTotal Weight 425 g Stir Rate 1250 RPMDI water needed 332.91 g Air Rate 0.00 LPMActual DI added 332 g Air Pressure 10 PSITemperature 35 °C 308 KActual Solids 20.04% Eh_Ag/AgCl 0.1889 V (Zhang 2004)Actual S/L ratio 25.06% ORP offsetRecorded ActualInitial NaOH dose 7.5 kg/t Update 216.7 217NaOH purity 98% Update 418.8 400NaOH needed 0.6502 g x (ORP Recorded) y (ORP Actual)Actual NaOH added g 216.7 217 Updatey = 0.9055x + 20.78R² = 1-600-400-2000200400600800-500 -300 -100 100 300 500ORP Calibration CurveSodium Hydroxide Dosage Formula Weight NaOH 39.997 g/mol 1.046 g/mLActual NaOH dosageOxidation NaOH Molarity 1.000 M 0 kg/t1.0 N NaOH 1.0 N NaOHTime (Hours) NaOH Weight (g) NaOH Dosed (g) Volume NaOH Dosed (mL) Total Volume NaOH Dosed (mL) NaOH added0 0 0 0.00 0.00 0 g solid1 0 0 0.00 0.00 0 moles2 0 0 0.00 0.00 0.00 mL sol'n4 0 0 0.00 0.00 0 moles8 0 0 0.00 0.0024 0 0 0.00 0.00 Oxidation0 molesTotal 0 0.00 0 g0 kg/tpH Conditioning1.0 N NaOHNaOH Weight (g) NaOH Dosed (g) pH Volume NaOH Dosed (mL) Total Volume NaOH Dosed (mL) NaOH addedStart 0 0 0.00 mL sol'nEnd 0 0.00 0.00 0.00 0 molesCyanidation Cyanidation ConditionspH ConditioningCyanide Dosage 10 kg/t Stir Rate 2000 RPM 0 molesEstimated Ore Charge 84.96 g Air Rate LPM 0 gCyanide Purity 94.9% Air Pressure PSI 0 kg/tCyanide Needed 0.8953 g Temperature 35 °C 308 KActual Cyanide Dosed 0.8953 g Eh_Ag/AgCl 0.1889 V(Zhang 2004)Cyanidation1.0 N NaOH 1.0 N NaOHTime (Hours) NaOH Weight (g) NaOH Dosed (g) Volume NaOH Dosed (mL) Total Volume NaOH Dosed (mL) NaOH added0 0 0 0.00 0.00 0.00 mL sol'n1 0 0 0.00 0.00 0 moles2 0 0 0.00 0.004 0 0 0.00 0.00 Cyanidation8 0 0 0.00 0.00 0 moles24 0 0 0.00 0.00 0 g0 kg/tTotal 0 0.00TotalEstimated Ore Charge 84.96 g Estimated Cyanide Dosage 10.00 kg/t 0 kg/t initialActual Ore Charge 84.96 g Actual Cyanide Dosage 10.00 kg/t0 Moles Added157    Initial Weight 1758.1 g Pre-24 Hour Weight 1837.1 g PLS Volume 535 mLInitial Solids 84.96 g Pre-24 Hour Solids 83.02 g PLS Weight 539.8 gInitial Solution 468.34 g Pre-24 Hour Solution 549.28 gPLS Density 1.009 g/mLInitial Solids % 15.36% 24 Hour Weight 1845.9 g24 Hour Solids 82.94 g Wash Volume 250 mL8 Hour Weight 1836.6 g 24 Hour Solution 558.16 g Wash Weight 248.6 g8 Hour Solids 84.29 g8 Hour Solution 547.51 g Final Solids % 12.94% Wash Density 0.994 g/mLRaw Data Update ConversionTime (Hours) pH ORP (RmV) ORP (mV) ORP (mV SHE) DO (mg/L) Temperature (°C) Temperature (K)0 11.87 162.1 168 356.1 4.8 35.4 308.61 11.86 121.1 130 318.7 4.9 35.6 308.82 11.84 111.3 122 310.1 5.0 35.4 308.64 11.79 112.1 122 310.8 5.1 35.4 308.68 11.73 104.0 115 303.3 5.0 35.5 308.724 11.22 105.7 116 305.2 5.1 35.2 308.4Time (Hours) Solution Aliquot (g) Solution Aliquot (L)Au (mg/L) Au w/ DF (mg/L) Au w/o DF (mg) Au w/ DF (mg)0 0.00 0.00 0.0000 0.00001 8.43 0.00836 1.66 1.80 0.0139 0.01512 8.57 0.00849 1.56 1.75 0.0133 0.01484 8.61 0.00853 1.50 1.73 0.0128 0.01488 8.74 0.00866 1.50 1.78 0.0130 0.015424 10.81 0.01071 1.48 1.76 0.0159 0.0189Wash 9.56 0.00961 0.10 0.1 0.0010 0.0010Total 0.0697 0.0800 mgTails Drying Tails DataBefore Before Au 2.47 g/mtFilter 0.94 g Funnels g Total S 1.02% (LECO)Towels 7.49 g Reactor 1205 gBag + Label 20.45 g Stirrer 43.12 gAfter AfterFilter 1.07 g Funnels gTowels 7.41 g Reactor 1204.8 gBag + Label 103.13 g Stirrer 43.11Spillage g Total Tails 82.94 gFunnels Residue 0 g Au in Tails 0.205 mgReactor Residue 0.2 g Solids Dissolved 2.02 gStirrer Residue 0.01 g Dissolution 2.38%Time (Hours) Total Weight (g) Solids Weight (g) Solution Weight (g) Solids Balance0 1758.10 84.96 468.34 15.36% 0.001248 1836.60 84.29 547.51 13.34% 0.0023 1837.10 83.02 549.28 13.13% 0.0024 1845.90 82.94 558.16 12.94% 0.00Total (after 8 hours) 20Washes 3Total Washes 60Water Added 79.17 gVolume/wash 1.32 gTime (Hours) Solution Weight (g) Probes Used Weight Added (g) New Weight (g) Solution Weight (g) Dilution Factor0 468.34 4 15.83 484.17 468.34 1.001 4 15.83 500.01 508.44 1.092 4 15.83 515.84 524.41 1.124 4 15.83 531.68 540.29 1.158 547.51 4 15.83 547.51 556.25 1.1924 558.16 4 1.91 558.16 558.16 1.19real Sol'n weight 558.16 1.19Notes:158  Total GoldSamples Volume (L) Au w/o (mg/L) Au w/o DF (mg)1 Hour Aliquot 0.008 1.66 0.0142 Hour Aliquot 0.008 1.56 0.0134 Hour Aliquot 0.009 1.50 0.0138 Hour Aliquot 0.009 1.50 0.013PLS 0.535 1.48 0.792Wash 0.250 0.10 0.025Solution Au 0.87 mgTails Au 0.20 mgTotal Au 1.07 mgw/o DF ErrorCalculated Head 12.65 g/mt -1.52 g/mtMass Balance w/o DFTotal Extraction 80.94%Modelled Extraction Au Concentration Au in Solution Extraction Extent of ReactionTime (Hours) Solution Volume (L) Au w/o DF (mg/L) Au w/ DF (mg/L) w/o DF (mg) w/ DF (mg) w/o DF (%) w/ DF (%) w/o DF (%) w/ DF (%)0 0.464 0.00 0.00 0.000 0.000 0.0% 0.0% 0% 0%1 0.504 1.66 1.80 0.975 0.889 90.8% 82.7% 112% 102%2 0.520 1.56 1.75 0.917 0.861 85.3% 80.2% 105% 99%4 0.535 1.50 1.73 0.881 0.853 82.0% 79.4% 101% 98%8 0.551 1.50 1.78 0.881 0.878 82.0% 81.7% 101% 101%24 0.553 1.48 1.76 0.870 0.870 80.9% 80.9% 100% 100%pre-filtration (no wash, no sampling)Actual 24 Hour vol: 0.588Error (incl. samples) 0.10 mLFormula Weight NaCN 49.0072 g/molPrimary Rxn AgNO3(aq) + 2NaCN(aq) = NaAg(CN)2(aq) + NaNO3(aq)Secondary Rxn AgNO3(aq) + KI(aq) = AgI(s) + KNO3(aq)Cyanide TitrationAgNO3 molarity 0.051 M AgNO3 start vol. 19.92 mL Initial conc. 1.83 g/L% KI indicator 5% w/w AgNO3 end vol. 23.21 mL NaCN conc. 0.75 g/LStirrer Residue Hours AgNO3 used 3.29 mL Conc. drop 1.08 g/LVolume Solution 25 mL Moles AgNO3 0.0002 SCN conc. mg/LConversion Rate 1.147 NaCN (SCN) 0.00 kg/tActual Volume 21.80 mL NaCN present 0.0003 moles in sample Initial NaCN level 10.00 kg/tOR NaCN final 4.91 kg/tSolution Weight g Initial NaCN 0.02 moles NaCN drop 5.09 kg/tPLS Density 1.009 g/mL Total NaCN (PLS) 0.01 moles in Sol'n NaCN (Au) 0.00 kg/tVolume Solution 0.00 mL NaCN drop 0.01 moles NaCN (non Au) 5.08 kg/t159  High Temperature Oxidation (Test F):   Alkaline Oxidation Ore Types MoistureTest F 165 0.1096 0.077Date 05-04-2017 167 0.1218 0.1662Ore Type 1 169 0.1392165 201 0.1192Moisture 7.70% 24-04-2017FileAssay Data Phases % Results UBC Proj 170403, 170410Au - Head FA 14.16 g/mt Actinolite 0 17-390-00174Ag - ICP-MS 9.50 g/mt Barite 6.2Fe - ICP-MS 1.79 % Goyazite 0Total S (LECO) 0.88 % Gypsum 0Ba - ICP-MS 20188.00 g/mt Hematite 0.3Al - ICP-MS 1.00 % Illite/Muscovite 2M1 1.3As - ICP-MS 172.00 g/mt Iron 0.4Hg - ICP-MS 10.64 g/mt Jarosite 0Kaolinite 0.8Reactor 1 Magnetite 0.1Reactor Weight 1185 g Marcasite 0Melanterite 0Needed 85 g Pyrite 0Ore Weighed 92.09 g Quartz 90.9Actual Ore Added 92.09 g Sulfur 0Solids 20% Total 100Dry Ore Added 85.00 g Oxidation ConditionsNaOH Initial Dosage 7.5 kg/tTotal Weight 425 g Stir Rate 1250 RPMDI water needed 332.91 g Air Rate 2.00 LPMActual DI added 332.5 g Air Pressure 10 PSITemperature 55 °C 328 KActual Solids 20.02% Eh_Ag/AgCl 0.1687 V (Zhang 2004)Actual S/L ratio 25.03% ORP offsetRecorded ActualInitial NaOH dose 7.5 kg/t Update 214.5 217NaOH purity 98% Update 426.8 400NaOH needed 0.6505 g x (ORP Recorded) y (ORP Actual)Actual NaOH added 0.6521 g 214.5 217 Updatey = 0.862x + 32.104R² = 1-600-400-2000200400600800-500 -300 -100 100 300 500ORP Calibration CurveInitial Weight 1620.9 g 8 Hour Weight 1620.2 g 24 Hour Weight 1614.1 gInitial Solids 85.00 g 8 Hour Solids 74.40 g 24 Hour Solids 63.26 gInitial Solution 350.90 g 8 Hour Solution 360.80 g 24 Hour Solution 365.84 gInitial Solids % 19.50% Final Solids % 14.74%Raw Data Update ConversionTime (Hours) pH ORP (RmV) ORP (mV) ORP (mV SHE) DO (mg/L) Temperature (°C) Temperature (K)0 12.58 -15.8 18 188.4 0.8 53.8 327.01 12.46 -29.1 7 177.2 0.7 53.5 326.72 12.43 -34.6 2 171.5 0.6 54.5 327.74 12.40 -47.4 -9 160.1 0.5 54.8 328.08 12.26 -34.8 2 170.7 0.6 55.1 328.324 12.96 -20.4 15 203.4 3.5 35.0 308.2Update Update Update UpdateRecorded Initial Wet Dry ActualTime (Hours) Solution Aliquot (g) Syringe Filter Wt (g)Syringe  Filter Wt (g) Syringe  Filter Wt (g) Solution Lost (g) Solids Lost (g) Solution Aliquot (g)01 8.69 2.60 2.76 2.62 0.15 0.02 8.542 8.46 2.60 2.76 2.60 0.17 0.00 8.294 8.68 2.60 2.71 2.60 0.12 0.00 8.568 8.34 0.00 10.42 8.3424 7.92 0.00 11.14 7.928.00 solution volume (mL)0.44 21.58Solid Filtration Total S (LECO) (%) g Sulphur Change8 Hours Initial Filter Weight (g) 0.92 -0.12%Filter After Drying (g) 11.34Filter After Sampling (g) 0.97 Total Sulphur LostSpillage (g) 0.260 gSample (g) 10.37 1.15% 0.11924 Hours Initial Filter Weight (g) 0.92Filter After Drying (g) 12.06Filter After Sampling (g) 0.96Spillage (g)Sample (g) 11.10 1.27% 0.141160    Time (Hours) Total Weight (g) Solids Weight (g) Solution Weight (g) Solids Balance0 1620.9 85.00 350.90 19.50% 0.001248 1620.20 74.40 360.80 17.10% 0.0024 1614.10 63.26 365.84 14.74% 0.00Total (after 8 hours) 20Washes 3Total Washes 60Water Added 9.90 gVolume/wash 0.16 mLTime (Hours) Solution Weight (g) Probes Used Water Added (g) New Weight (g) Solution Weight (g) Dilution Factor0 350.90 4 1.98 352.88 350.90 1.0001 4 1.98 354.86 363.55 1.0362 4 1.98 356.84 365.30 1.0414 4 1.98 358.82 367.50 1.0478 360.80 4 1.98 360.80 369.14 1.05224 365.84 4 -3.30 365.84 373.76 1.065real Sol'n weight 373.76 1.065Time (Hours) Solution Aliquot (g) Solution Aliquot (L)S conc. (g/L) S conc. w/ DF (g/L)S w/o DF (g) S w/ DF (g)0 0 0 01 8.69 0.00877 0.095 0.098 0.000833 0.0008632 8.46 0.00854 0.103 0.107 0.000879 0.0009154 8.68 0.00876 0.117 0.123 0.001025 0.0010738 8.34 0.00842 0.117 0.123 0.000985 0.00103624 7.92 0.00799 0.117 0.125 0.000935 0.000996S in Samples (g) 0.00466 0.00488Sample Density Formula Weight SO4 96.060.99 g/mL S 32.06Aqueous ConcentrationsTime (Hours) Aluminium (mg/L) Calcium (mg/L) Iron (mg/L) Magnesium (mg/L) Sulphur (mg/L) Sulphate (mg/L)0 0.00 0 0.00 0.00 0 0 w/o DF1 6.0 93.0 0.0 0.0 95.0 284.62 6.2 90.0 0.0 0.0 103.0 308.64 5.8 80.8 0.0 0.0 117.0 350.68 5.2 13.1 0.0 0.0 117.0 350.624 4.9 4.1 0.0 0.0 117.0 350.6Time (Hours) Aluminium (mg/L) Calcium (mg/L) Iron (mg/L) Magnesium (mg/L) Sulphur (mg/L) Sulphate (mg/L)0 0.0 0.0 0.0 0.0 0.0 0.0 w/ DF1 6.2 96.4 0.0 0.0 98.4 294.92 6.5 93.7 0.0 0.0 107.2 321.34 6.1 84.6 0.0 0.0 122.5 367.18 5.5 13.8 0.0 0.0 123.1 368.824 5.2 4.4 0.0 0.0 124.6 373.4Total Sulphur Sulphur Concentration Sulphur Content Oxidation Extent of ReactionTime (Hours) Solution Volume (L) w/o DF (g/L) w/ DF (g/L) w/o DF (g) SO4 w/ DF (g) w/o DF (%) w/ DF (%) w/o DF (%) w/ DF (%)0 0.354 0.000 0.000 0.0000 0.0000 0.00% 0.00% 0% 0%1 0.367 0.095 0.098 0.0386 0.0380 3.77% 3.72% 81% 80%2 0.369 0.103 0.107 0.0418 0.0414 4.09% 4.05% 88% 87%4 0.371 0.117 0.123 0.0475 0.0473 4.65% 4.63% 100% 100%8 0.372 0.117 0.123 0.0475 0.0475 4.65% 4.65% 100% 100%Method 1 Using LECO tests 8 Hour BasisOxidation (Residue) Total S (LECO) (%) Actual Oxidation Sol'n S Content 0.047 g8 Hours 1.15% 4.65% Solid S Content 0.976 gTotal S (8 hr) 1.023 gS content (%) 1.20% calculatedMethod 2 Using ICP-AES Error 0.32%Oxidation (Solution) w/o DF w/ DF Error Calc. S Content 12.04 kg/t8 Hours 4.65% 4.65% 0.00% ΔS (8 hours) 0.56 kg/t "removed"161    Sodium Hydroxide Dosage Formula Weight NaOH 39.997 g/mol 1.046 g/mLActual NaOH dosageOxidation NaOH Molarity 1.000 M 7.52 kg/t1.0 N NaOH 1.0 N NaOHTime (Hours) NaOH Weight (g) NaOH Dosed (g) Volume NaOH Dosed (mL) Total Volume NaOH Dosed (mL) NaOH added0 0 0.00 0.00 0.64 g solid1 0 0.00 0.00 0.02 moles2 0 0.00 0.00 0.00 mL sol'n4 0 0.00 0.00 0 moles8 0 0.00 0.0024 0 0.00 0.00 Oxidation0.02 molesTotal 0 0.00 0.64 g7.52 kg/tpH Conditioning1.0 N NaOHNaOH Weight (g) NaOH Dosed (g) pH Volume NaOH Dosed (mL) Total Volume NaOH Dosed (mL) NaOH addedStart 0 0.00 mL sol'nEnd 0.00 0.00 0.00 0 molesCyanidation Cyanidation ConditionspH ConditioningCyanide Dosage 10 kg/t Stir Rate 2000 RPM 0 molesEstimated Ore Charge 65.67 g Air Rate LPM 0 gCyanide Purity 94.9% Air Pressure PSI 0 kg/tCyanide Needed 0.6920 g Temperature 35 °C 308 KActual Cyanide Dosed 0.6926 g Eh_Ag/AgCl 0.1889 V(Zhang 2004)Cyanidation1.0 N NaOH 1.0 N NaOHTime (Hours) NaOH Weight (g) NaOH Dosed (g) Volume NaOH Dosed (mL) Total Volume NaOH Dosed (mL) NaOH added0 0 0.00 0.00 0.00 mL sol'n1 0 0.00 0.00 0 moles2 0 0.00 0.004 0 0.00 0.00 Cyanidation8 0 0.00 0.00 0 moles24 0 0.00 0.00 0 g0 kg/tTotal 0 0.00TotalEstimated Ore Charge 65.67 g Estimated Cyanide Dosage 10.01 kg/t 7.52 kg/t initialActual Ore Charge 63.26 g Actual Cyanide Dosage 10.39 kg/t0.02 Moles AddedInitial Weight 1630.7 g Pre-24 Hour Weight 1666.6 g PLS Volume 415 mLInitial Solids 63.26 g Pre-24 Hour Solids 62.84 g PLS Weight 411.3 gInitial Solution 382.44 g Pre-24 Hour Solution 418.76 gPLS Density 0.991 g/mLInitial Solids % 14.19% 24 Hour Weight 1674.1 g24 Hour Solids 62.82 g Wash Volume 260 mL8 Hour Weight 1666 g 24 Hour Solution 426.28 g Wash Weight 256.6 g8 Hour Solids 63.11 g 11.288 Hour Solution 417.89 g Final Solids % 12.84% Wash Density 0.987 g/mLRaw Data Update ConversionTime (Hours) pH ORP (RmV) ORP (mV) ORP (mV SHE) DO (mg/L) Temperature (°C) Temperature (K)0 12.81 136.3 150 338.0 3.6 35.5 308.71 12.79 126.5 141 329.3 3.7 35.7 308.92 12.73 130.5 145 333.0 3.3 35.5 308.74 12.55 115.5 132 320.0 3.3 35.6 308.88 12.16 123.7 139 327.1 3.2 35.5 308.724 11.14 115.3 131 320.0 3.5 35.4 308.6Time (Hours) Solution Aliquot (g) Solution Aliquot (L)Au (mg/L) Au w/ DF (mg/L) Au w/o DF (mg) Au w/ DF (mg)0 0.00 0.00 0.0000 0.00001 8.50 0.00858 1.60 1.69 0.0137 0.01452 8.49 0.00857 1.58 1.70 0.0135 0.01464 8.44 0.00852 1.54 1.69 0.0131 0.01448 8.21 0.00828 1.46 1.63 0.0121 0.013524 9.66 0.00975 1.48 1.65 0.0144 0.0161Wash 9.09 0.00921 0.07 0.07 0.0006 0.0006Total 0.0675 0.0737 mgTails Drying Tails DataBefore Before Au 2.32 g/mtFilter 0.92 g Funnels g Total S 1.07% (LECO)Towels 10.44 g Reactor 1185.6 gBag + Label 19.07 g Stirrer 43.12 gAfter AfterFilter 1.15 g Funnels gTowels 10.32 g Reactor 1185.3 gBag + Label 81.46 g Stirrer 43.1Spillage g Total Tails 62.82 gFunnels Residue 0 g Au in Tails 0.146 mgReactor Residue 0.3 g Solids Dissolved 0.62 gStirrer Residue 0.02 g Dissolution 0.73%162    Time (Hours) Total Weight (g) Solids Weight (g) Solution Weight (g) Solids Balance0 1630.70 63.26 382.44 14.19% 0.001248 1666.00 63.11 417.89 13.12% 0.0023 1666.60 62.84 418.76 13.05% 0.0024 1674.10 62.82 426.28 12.84% 0.00Total (after 8 hours) 20Washes 3Total Washes 60Water Added 35.45 gVolume/wash 0.59 gTime (Hours) Solution Weight (g) Probes Used Weight Added (g) New Weight (g) Solution Weight (g) Dilution Factor0 382.44 4 7.09 389.53 382 1.001 4 7.09 396.62 405 1.062 4 7.09 403.71 412 1.084 4 7.09 410.80 419 1.108 417.89 4 7.09 417.89 426 1.1124 426.28 4 0.18 426.28 426 1.11real Sol'n weight 426 1.11Notes:Total GoldSamples Volume (L) Au w/o (mg/L) Au w/o DF (mg)1 Hour Aliquot 0.009 1.60 0.0142 Hour Aliquot 0.009 1.58 0.0144 Hour Aliquot 0.009 1.54 0.0138 Hour Aliquot 0.008 1.46 0.012PLS 0.415 1.48 0.614Wash 0.260 0.07 0.018Solution Au 0.685 mgTails Au 0.146 mgTotal Au 0.831 mgw/o DF ErrorCalculated Head 13.13 g/mt -1.03 g/mtMass Balance w/o DFTotal Extraction 82.45%Modelled Extraction Au Concentration Au in Solution Extraction Extent of ReactionTime (Hours) Solution Volume (L) Au w/o DF (mg/L) Au w/ DF (mg/L) w/o DF (mg) w/ DF (mg) w/o DF (%) w/ DF (%) w/o DF (%) w/ DF (%)0 0.386 0.00 0.00 0.000 0.000 0.0% 0.0% 0% 0%1 0.409 1.60 1.69 0.740 0.704 89.1% 84.7% 108% 103%2 0.416 1.58 1.70 0.731 0.707 88.0% 85.1% 107% 103%4 0.423 1.54 1.69 0.713 0.701 85.8% 84.4% 104% 102%8 0.430 1.46 1.63 0.676 0.675 81.3% 81.3% 99% 99%24 0.430 1.48 1.65 0.685 0.685 82.5% 82.5% 100% 100%pre-filtration (no wash, no sampling)Actual 24 Hour vol: 0.463 LError (incl. samples) -1.01 mLFormula Weight NaCN 49.0072 g/molPrimary Rxn AgNO3(aq) + 2NaCN(aq) = NaAg(CN)2(aq) + NaNO3(aq)Secondary Rxn AgNO3(aq) + KI(aq) = AgI(s) + KNO3(aq)Cyanide TitrationAgNO3 molarity 0.051 M AgNO3 start vol. 3.61 mL Initial conc. 1.70 g/L% KI indicator 5% w/w AgNO3 end vol. 5.55 mL NaCN conc. 0.39 g/LStirrer Residue Hours AgNO3 used 1.94 mL Conc. drop 1.31 g/LVolume Solution mL Moles AgNO3 0.0001 SCN conc. mg/LConversion Rate 1.147 NaCN (SCN) 0.00 kg/tActual Volume 0 mL NaCN present 0.0002 moles in sample Initial NaCN level 10.39 kg/tOR NaCN final 2.65 kg/tSolution Weight 24.68 g Initial NaCN 0.0134 moles NaCN drop 7.74 kg/tPLS Density 0.991 g/mL Total NaCN (PLS) 0.0034 moles in Sol'n NaCN (Au) 0.00 kg/tVolume Solution 24.90 mL NaCN drop 0.0100 moles NaCN (non Au) 7.74 kg/t

Cite

Citation Scheme:

        

Citations by CSL (citeproc-js)

Usage Statistics

Share

Embed

Customize your widget with the following options, then copy and paste the code below into the HTML of your page to embed this item in your website.
                        
                            <div id="ubcOpenCollectionsWidgetDisplay">
                            <script id="ubcOpenCollectionsWidget"
                            src="{[{embed.src}]}"
                            data-item="{[{embed.item}]}"
                            data-collection="{[{embed.collection}]}"
                            data-metadata="{[{embed.showMetadata}]}"
                            data-width="{[{embed.width}]}"
                            async >
                            </script>
                            </div>
                        
                    
IIIF logo Our image viewer uses the IIIF 2.0 standard. To load this item in other compatible viewers, use this url:
https://iiif.library.ubc.ca/presentation/dsp.24.1-0366302/manifest

Comment

Related Items