Applied Science, Faculty of
Engineering, School of (Okanagan)
DSpace
UBCO
Mackowiak, Kristian David
Doctor of Philosophy - PhD
University of British Columbia
The objective of the current research was the extraction and separation of silver and tellurium from an industrial slag by-product. The recovery of these elements presented an opportunity to extract additional value from the slag, as well as produce tellurium, an important metal in solar panel production. A 2N purity tellurium powder was produced using a three-stage process. The process first involved a pyrometallurgical processing step, which served to separate the silver from tellurium. A thermodynamic model was first produced in HSC 9.1 to determine the effects of temperature and various reagents on the pyrometallurgical system. Using the model results, a central composite design (CCD) methodology was utilized to determine the experimental tests to conduct. The results of the CCD testing were used to generate full quadratic statistical models using MATLAB®. From the experimental testing, the silver recovery during the pyrometallurgical step was 92% to the metal phase and 88% tellurium recovery to the slag phase. The remaining pyrometallurgically processed slag from the previous step was then crushed and prepared for leaching. The lixiviant used was deionized water, with the slurry consisting of 1 to 30% slag by weight. Approximately 30% of the Te in the slag would instantaneously leach, which was attributed to sodium tellurite dissolution. Following 24 hours of leaching, additional tellurium would leach into solution. It was found that a lower solids percentage in the slurry yielded over 60% Te recovery. The addition of sodium hydroxide or sodium sulphide also led to increases in the Te recovery. The resulting tellurium leach solution could then be purified in one of two ways. First, electrowinning of the solution could be conducted, which led to the production of the 2N purity Te powders. Recoveries typically around 70 to 90% were achieved in a single electrowinning step. It was found that using the leach solution containing sodium sulphide would lead to the highest Te recovery, at 91%, with a purity of nearly 3N. Combining all three steps, the overall Te recovery of the system was found to be 46%.
Recovery and Production ofHigh-Purity Tellurium from SilverSoda SlagbyKristian David MackowiakBASc, Queen’s University, 2014MASc, Queen’s University, 2016A DISSERTATION SUBMITTED IN PARTIAL FULFILLMENT OFTHE REQUIREMENTS FOR THE DEGREE OFDOCTOR OF PHILOSOPHYinThe College of Graduate Studies(Mechanical Engineering)THE UNIVERSITY OF BRITISH COLUMBIA(Okanagan)May 2021© Kristian David Mackowiak, 2021The following individuals certify that they have read and recommend to the College ofGraduate Studies for acceptance, a dissertation entitled:Recovery and Production of High-Purity Tellurium from Silver SodaSlagsubmitted by Kristian David Mackowiak in partial fulfilment of the requirements ofthe degree of Doctor of Philosophy.Dr. Lukas Bichler, School of EngineeringSupervisorDr. Dimitry Sediako, School of EngineeringSupervisory Committee MemberDr. Leili Tafaghodi, Department of Materials Science and Engineering, McMaster UniversitySupervisory Committee MemberDr. Cigdem Eskicioglu, School of EngineeringUniversity ExaminerDr. Mansoor Barati, Department of Materials Science and Engineering, University of TorontoExternal ExamineriiAbstractThe objective of the current research was the extraction and separation of silver andtellurium from an industrial slag by-product. The recovery of these elements presentedan opportunity to extract additional value from the slag, as well as produce tellurium, animportant metal in solar panel production.A 2N purity tellurium powder was produced using a three-stage process. The processfirst involved a pyrometallurgical processing step, which served to separate the silver fromtellurium. A thermodynamic model was first produced in HSC 9.1 to determine the effects oftemperature and various reagents on the pyrometallurgical system. Using the model results,a central composite design (CCD) methodology was utilized to determine the experimentaltests to conduct. The results of the CCD testing were used to generate full quadraticstatistical models using MATLAB®. From the experimental testing, the silver recoveryduring the pyrometallurgical step was 92% to the metal phase and 88% tellurium recoveryto the slag phase.The remaining pyrometallurgically processed slag from the previous step was thencrushed and prepared for leaching. The lixiviant used was deionized water, with the slurryconsisting of 1 to 30% slag by weight. Approximately 30% of the Te in the slag would instan-taneously leach, which was attributed to sodium tellurite dissolution. Following 24 hoursof leaching, additional tellurium would leach into solution. It was found that a lower solidspercentage in the slurry yielded over 60% Te recovery. The addition of sodium hydroxideor sodium sulphide also led to increases in the Te recovery.The resulting tellurium leach solution could then be purified in one of two ways. First,electrowinning of the solution could be conducted, which led to the production of the2N purity Te powders. Recoveries typically around 70 to 90% were achieved in a singleelectrowinning step. It was found that using the leach solution containing sodium sulphidewould lead to the highest Te recovery, at 91%, with a purity of nearly 3N.Combining all three steps, the overall Te recovery of the system was found to be 46%.iiiLay SummaryThe objective of this work was to investigate the separation and purification of tellurium,an important metal, from a mining by-product. Tellurium (Te) is a critical element in theproduction of renewable energy sources, primarily solar panels, which is the end use ofapproximately 40% of all Te produced worldwide. As such, the recovery of this element iscritical for the continued shift from non-renewable to renewable energy sources.The tellurium in this research originated from a by-product slag from lead production.Since this slag is not currently processed by the company to collect the Te, the objective ofthis research was to develop of a process to recover the Te and other economically valuableelements from the slag.This dissertation outlines the development and results of a three-stage process, which isable to recover the tellurium from the slag and purify it to 99% Te. This is an acceptablelevel of purity for further purification before the Te can be used in solar or other applications.ivPrefaceThe following copyright materials were used with permission:Figure A.1 Reprinted with permission from Springer Nature Customer Service Centre GmbH:Springer Journal of Phase Equilibria. The Bi-Pb (Bismuth-Lead) System, N. A.Gokcen. Copyright 1992.Figure 2.1 Reprinted with permission from Springer Nature Customer Service Centre GmbH:Springer Metallurgical and Materials Transactions B. Thermodynamics of re-moving selenium and tellurium from liquid copper by sodium carbonate slags, Kojo,Taskinen, and Lilius. Copyright 1985.Figure 2.2 Reprinted with permission from Elsevier Elsevier Materials Chemistry andPhysics. The microstructure, liquidus projection and thermodynamic modeling ofthermoelectric Ag–Pb–Te system, Wu, Foo, Gierlotka, Chen, and Snyder. Copyright2013.Figure A.2 Reprinted with permission from Springer Nature Customer Service Centre GmbH:Springer Bulletin of Alloy Phase Diagrams. The Ag-Pb (Silver-Lead) System,Karakaya and Thompson. Copyright 1987.Figure A.3 Reprinted with permission from Springer Nature Customer Service Centre GmbH:Springer Bulletin of Alloy Phase Diagrams. The Pb-Te (Lead-Tellurium) Sys-tem, Lin, Hsieh, Sharma, and Chang. Copyright 1989.Figure A.4 Reprinted with permission from Springer Nature Customer Service Centre GmbH:Springer Bulletin of Alloy Phase Diagrams. The Pb-Te (Lead-Tellurium) Sys-tem, Lin, Hsieh, Sharma, and Chang. Copyright 1989.Figure A.5 Reprinted with permission from Springer Nature Customer Service Centre GmbH:Springer Journal of Phase Equilibria. The Na-Te (Sodium-Tellurium) System,Pelton and Petric. Copyright 1990.vPrefaceFigure A.6 Reprinted with permission from Springer Nature Customer Service Centre GmbH:Springer Journal of Phase Equilibria. The Ag-Te (Silver-Tellurium) System,Karakaya and Thompson. Copyright 1991.Figure A.7 Reprinted with permission from Springer Nature Customer Service Centre GmbH:Springer Journal of Phase Equilibria. The Ag-Te (Silver-Tellurium) System,Karakaya and Thompson. Copyright 1991.viContentsAbstract . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . iiiLay Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . ivPreface . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . vContents . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . viiList of Tables . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . xiiiList of Figures . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . xvList of Abbreviations . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .xxxiList of Symbols . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .xxxiiList of Chemical Formulae . . . . . . . . . . . . . . . . . . . . . . . . . . . . .xxxiiiGlossary of Notation . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .xxxvAcknowledgements . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .xxxviChapter 1: Introduction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 11.1 General Overview . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 11.2 Tellurium . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 11.2.1 Uses . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 11.2.2 Economics . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 21.2.3 Sources . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 31.2.4 Current Processing Methods . . . . . . . . . . . . . . . . . . . . . . 31.3 Research Scope and Objectives . . . . . . . . . . . . . . . . . . . . . . . . . 41.4 Chapter Organization . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 4viiCONTENTSChapter 2: Literature Review . . . . . . . . . . . . . . . . . . . . . . . . . . . 62.1 Overview . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 62.2 Pyrometallurgical Processing . . . . . . . . . . . . . . . . . . . . . . . . . . 62.2.1 Thermodynamics . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 132.2.2 Pyrometallurgical Summary . . . . . . . . . . . . . . . . . . . . . . . 142.3 Hydrometallurgical Processing . . . . . . . . . . . . . . . . . . . . . . . . . 142.3.1 Acidic Leaching . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 152.3.2 Alkaline Leaching . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 182.3.3 Hydrometallurgical Summary . . . . . . . . . . . . . . . . . . . . . . 212.4 Te Recovery . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 222.4.1 Electrowinning . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 222.4.2 Copper Telluride (Cu2Te) Cementation . . . . . . . . . . . . . . . . 262.4.3 Electrowinning Summary . . . . . . . . . . . . . . . . . . . . . . . . 282.5 State of Current Literature . . . . . . . . . . . . . . . . . . . . . . . . . . . 28Chapter 3: Experimental Methods . . . . . . . . . . . . . . . . . . . . . . . . 303.1 Chemicals . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 303.2 Slags . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 303.3 Variables Investigated . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 333.3.1 Pyrometallurgical Variables . . . . . . . . . . . . . . . . . . . . . . . 333.3.2 Hydrometallurgical Variables . . . . . . . . . . . . . . . . . . . . . . 353.3.3 Electrowinning Variables . . . . . . . . . . . . . . . . . . . . . . . . . 363.3.4 Tellurium Dioxide Precipitation . . . . . . . . . . . . . . . . . . . . . 373.4 Experimental Methods . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 373.4.1 Pyrometallurgical Tests . . . . . . . . . . . . . . . . . . . . . . . . . 373.4.2 Hydrometallurgical Tests . . . . . . . . . . . . . . . . . . . . . . . . 403.4.3 Electrowinning . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 423.4.4 Tellurium Dioxide Precipitation . . . . . . . . . . . . . . . . . . . . . 433.5 Analytical Methods . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 443.5.1 Scanning Electron Microscope (SEM) . . . . . . . . . . . . . . . . . 443.5.2 Inductively Coupled Plasma Spectrometry (ICP-AES / ICP-MS) . . 453.5.3 X-Ray Diffraction (XRD) . . . . . . . . . . . . . . . . . . . . . . . . 453.5.4 Metallography . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 453.6 Statistical Analysis . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 463.6.1 Central Composite Design . . . . . . . . . . . . . . . . . . . . . . . . 463.6.2 Statistical Modelling . . . . . . . . . . . . . . . . . . . . . . . . . . . 48viiiCONTENTS3.7 Safety . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 49Chapter 4: Thermodynamics . . . . . . . . . . . . . . . . . . . . . . . . . . . . 514.1 Thermodynamic Considerations . . . . . . . . . . . . . . . . . . . . . . . . . 514.2 Model Limitations . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 514.3 Species and Phases Considered . . . . . . . . . . . . . . . . . . . . . . . . . 514.4 Activity Coefficients . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 524.5 Variables Investigated . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 544.6 Effect of Temperature . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 544.7 Effect of Fluxing Reagents . . . . . . . . . . . . . . . . . . . . . . . . . . . . 604.7.1 Effect of Silicon Dioxide . . . . . . . . . . . . . . . . . . . . . . . . . 604.7.2 Effect of Sodium Carbonate . . . . . . . . . . . . . . . . . . . . . . . 634.7.3 Effect of Lead(II) Oxide . . . . . . . . . . . . . . . . . . . . . . . . . 664.7.4 Effect of Lead(II) Oxide and Sodium Carbonate . . . . . . . . . . . 694.8 Effect of Initial Oxygen . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 724.8.1 Element Distribution at 1000 ◦C . . . . . . . . . . . . . . . . . . . . 764.9 Effect of Pre-Roasting . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 824.10 Leaching - Eh-pH Diagram . . . . . . . . . . . . . . . . . . . . . . . . . . . 874.11 Tellurium Dioxide Precipitation . . . . . . . . . . . . . . . . . . . . . . . . . 88Chapter 5: Results and Discussion - Pyrometallurgy . . . . . . . . . . . . . 905.1 Preliminary Results . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 905.1.1 Effect of Silica Addition . . . . . . . . . . . . . . . . . . . . . . . . . 905.1.2 Effect of Lead Oxide Addition . . . . . . . . . . . . . . . . . . . . . 935.1.3 Effect of Sodium Carbonate Addition . . . . . . . . . . . . . . . . . 945.1.4 Effect of Temperature and Time . . . . . . . . . . . . . . . . . . . . 945.1.5 Effect of Pre-Roast . . . . . . . . . . . . . . . . . . . . . . . . . . . . 955.2 Central Composite Design Results . . . . . . . . . . . . . . . . . . . . . . . 955.2.1 Center Point Results . . . . . . . . . . . . . . . . . . . . . . . . . . . 965.2.2 Statistical Modelling Results . . . . . . . . . . . . . . . . . . . . . . 1005.2.3 Distribution of Other Elements . . . . . . . . . . . . . . . . . . . . . 1175.2.4 Model Applications and Optimization . . . . . . . . . . . . . . . . . 1205.3 Microstructure Characterization . . . . . . . . . . . . . . . . . . . . . . . . 1215.4 Thermodynamic Model Comparison . . . . . . . . . . . . . . . . . . . . . . 1255.5 Processed Slag Analysis and Proposed Reactions . . . . . . . . . . . . . . . 1275.5.1 Effect of Temperature . . . . . . . . . . . . . . . . . . . . . . . . . . 1285.5.2 Effect of Time . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 129ixCONTENTS5.5.3 Effect of PbO Addition . . . . . . . . . . . . . . . . . . . . . . . . . 1315.5.4 Effect of Na2CO3 Addition . . . . . . . . . . . . . . . . . . . . . . . 1335.5.5 Comparison to Thermodynamics . . . . . . . . . . . . . . . . . . . . 1345.5.6 Combined Leach Input Slag . . . . . . . . . . . . . . . . . . . . . . . 1375.6 Pyrometallurgical Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . 138Chapter 6: Results and Discussion - Hydrometallurgy . . . . . . . . . . . . 1406.1 Preliminary Leaching Results . . . . . . . . . . . . . . . . . . . . . . . . . . 1406.2 Main Series Leaching Results . . . . . . . . . . . . . . . . . . . . . . . . . . 1426.2.1 Effect of Pre-Leach Roasting . . . . . . . . . . . . . . . . . . . . . . 1436.2.2 Solution pH . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1466.2.3 Tellurium Leaching Behaviour . . . . . . . . . . . . . . . . . . . . . 1506.2.4 Tellurium PLS . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1646.3 Leaching Residue Analysis . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1656.4 Leaching Summary . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1686.5 Te PLS Sodium Sulphide Cleaning . . . . . . . . . . . . . . . . . . . . . . . 170Chapter 7: Results and Discussion - Tellurium Extraction . . . . . . . . . . 1737.1 Preliminary Electrowinning Results . . . . . . . . . . . . . . . . . . . . . . . 1737.2 Electrowinning Tellurium Recovery . . . . . . . . . . . . . . . . . . . . . . . 1767.3 Current Efficiency . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1827.4 Tellurium Product . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1877.4.1 Te Grade . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1877.4.2 Impurity Analysis . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1907.4.3 XRD Analysis . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1927.5 SEM Images . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 1947.6 Electrowinning Results Summary . . . . . . . . . . . . . . . . . . . . . . . . 2057.7 Tellurium Dioxide Precipitation . . . . . . . . . . . . . . . . . . . . . . . . . 2077.8 Tellurium Dioxide SEM Images . . . . . . . . . . . . . . . . . . . . . . . . . 2157.9 Tellurium Dioxide Precipitation Summary . . . . . . . . . . . . . . . . . . . 224Chapter 8: Results and Discussion - Flowsheet . . . . . . . . . . . . . . . . . 2258.1 Process Flowsheet . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2258.2 Suggested Flowsheet Improvements . . . . . . . . . . . . . . . . . . . . . . . 226Chapter 9: Conclusion and Recommendations for Future Work . . . . . . . 2279.1 Thermodynamic Model . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 227xCONTENTS9.2 Pyrometallurgical Testing . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2279.2.1 Central Composite Design Results . . . . . . . . . . . . . . . . . . . 2289.2.2 XRD Results . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2299.2.3 Comparison of Experimental Results to Thermodynamics . . . . . . 2319.3 Hydrometallurgical Leaching . . . . . . . . . . . . . . . . . . . . . . . . . . 2319.3.1 Effect of Leaching Parameters on Te Recovery . . . . . . . . . . . . 2319.3.2 Leach Solution Cleaning . . . . . . . . . . . . . . . . . . . . . . . . . 2339.4 Tellurium Extraction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2349.4.1 Electrowinning . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 2349.4.2 Tellurium Dioxide Precipitation . . . . . . . . . . . . . . . . . . . . . 2369.5 Future Work . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 236Bibliography . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 237Appendix . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 244Appendix A: Phase Diagrams . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 245A.1 Phase Diagrams . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 245A.1.1 Bismuth-Lead System . . . . . . . . . . . . . . . . . . . . . . . . . . 245A.1.2 Silver-Lead System . . . . . . . . . . . . . . . . . . . . . . . . . . . . 246A.1.3 Lead-Tellurium System . . . . . . . . . . . . . . . . . . . . . . . . . 247A.1.4 Sodium-Tellurium System . . . . . . . . . . . . . . . . . . . . . . . . 249A.1.5 Silver-Tellurium System . . . . . . . . . . . . . . . . . . . . . . . . . 250Appendix B: Schematic Drawings . . . . . . . . . . . . . . . . . . . . . . . . . . . 253B.1 Experimental Design Drawings . . . . . . . . . . . . . . . . . . . . . . . . . 253Appendix C: Supplemental Thermodynamic Modelling Results . . . . . . . . . . 256C.1 Effect of Temperature . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 256C.2 Effect of Fluxing Reagents . . . . . . . . . . . . . . . . . . . . . . . . . . . . 261Appendix D: Supplemental Pyrometallurgical Testing Results . . . . . . . . . . . 265D.1 Pyrometallurgical Experimental Results . . . . . . . . . . . . . . . . . . . . 265Appendix E: Additional Leaching Figures . . . . . . . . . . . . . . . . . . . . . . 267E.0.1 Selenium Leaching Behaviour . . . . . . . . . . . . . . . . . . . . . . 267E.0.2 Silicon Leaching Behaviour . . . . . . . . . . . . . . . . . . . . . . . 276E.0.3 Aluminum Leaching Behaviour . . . . . . . . . . . . . . . . . . . . . 286E.0.4 Arsenic Leaching Behaviour . . . . . . . . . . . . . . . . . . . . . . . 295E.1 Other Element Behaviour . . . . . . . . . . . . . . . . . . . . . . . . . . . . 304Appendix F: Full Quadratic Models . . . . . . . . . . . . . . . . . . . . . . . . . 306F.1 Thermodynamic Model Fitting . . . . . . . . . . . . . . . . . . . . . . . . . 306xiCONTENTSF.2 Pyrometallurgical Models . . . . . . . . . . . . . . . . . . . . . . . . . . . . 308xiiList of TablesTable 3.1 Chemicals used during experimentation. . . . . . . . . . . . . . . . . 30Table 3.2 Composition (in wt. %) of slags examined in this thesis. . . . . . . . 31Table 3.3 Preliminary pyrometallurgical testing variables and values. . . . . . . 34Table 3.4 Input variables and values examined during pyrometallurgical tests. 34Table 3.5 Preliminary leaching variables and values. . . . . . . . . . . . . . . . 35Table 3.6 Input variables and values examined during leaching tests. . . . . . . 36Table 3.7 Preliminary electrowinning variables and values. . . . . . . . . . . . 36Table 3.8 Input variables and values examined during electrowinning tests. . . 37Table 3.9 Matrix for CCD design generated using MATLAB®, indicating testnumber. Test numbers with a C are center points. . . . . . . . . . . 47Table 3.10 Values used in CCD testing. . . . . . . . . . . . . . . . . . . . . . . . 48Table 4.1 Species and phases considered in the thermodynamic model. . . . . . 52Table 5.1 Comparison of pyro and pre-roast and pyro on several response vari-ables in the metal phase. The values in brackets indicate the standarddeviation (σ), three replicates were conducted per test. . . . . . . . . 95Table 5.2 Pyrometallurgical CCD testing statistical model output. . . . . . . . 102Table 6.1 PLS compositions. . . . . . . . . . . . . . . . . . . . . . . . . . . . . 165Table 6.2 Leaching summary. . . . . . . . . . . . . . . . . . . . . . . . . . . . . 169Table 7.1 Parameters and R2 value for curve fit used for the preliminary elec-trowinning tests. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 174Table 7.2 Tellurium grades obtained via ICP-MS for preliminary electrowinningtests. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 175Table 7.3 Parameters and R2 value for curve fit used. . . . . . . . . . . . . . . 177Table 7.4 Tellurium grade and standard deviations for products of each elec-trowinning test, obtained via ICP-MS. . . . . . . . . . . . . . . . . . 188xiiiLIST OF TABLESTable D.1 Experimentally obtained values for the recoveries of tellurium andsilver, and the total impurity grade, used as input to the CCD model. 266Table F.1 Full quadratic model output from MATLAB® for the thermodynamicCCD model showing the coefficients for silver recovery to the metalphase. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 306Table F.2 Full quadratic model output from MATLAB® for the thermodynamicCCD model showing the coefficients for tellurium recovery to the slagphase. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 307Table F.3 Full quadratic model output from MATLAB® for the pyrometallur-gical testing model showing the coefficients for tellurium recovery tothe slag phase. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 308Table F.4 Full quadratic model output from MATLAB® for the pyrometallur-gical testing model showing the coefficients for tellurium recovery tothe metal phase. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 309Table F.5 Full quadratic model output from MATLAB® for the pyrometallur-gical testing model showing the coefficients for silver recovery to themetal phase. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 310Table F.6 Full quadratic model output from MATLAB® for the pyrometallurgi-cal testing model showing the coefficients for the total impurity gradeof the metal phase. . . . . . . . . . . . . . . . . . . . . . . . . . . . . 311xivList of FiguresFigure 1.1 Price of tellurium in 2020 USD per kg, from USGS [6]. . . . . . . . 2Figure 1.2 World production values from USGS, values after 2003 excluded USto avoid disclosing company data [6, 8]. . . . . . . . . . . . . . . . . 3Figure 2.1 Equilibrium distribution of selenium and tellurium between copperand soda slag [13]. Reprinted with permission from Springer NatureCustomer Service Centre GmbH: Springer Metallurgical and Ma-terials Transactions B. Thermodynamics of removing seleniumand tellurium from liquid copper by sodium carbonate slags, Kojo,Taskinen, and Lilius. Copyright 1985. . . . . . . . . . . . . . . . . . 8Figure 2.2 The bismuth-lead (Bi-Pb) system [23]. Reprinted with permissionfrom Elsevier Elsevier Materials Chemistry and Physics. Themicrostructure, liquidus projection and thermodynamic modeling ofthermoelectric Ag–Pb–Te system, Wu, Foo, Gierlotka, Chen, andSnyder. Copyright 2013. . . . . . . . . . . . . . . . . . . . . . . . . 14Figure 3.1 Optical image of the as-received slag B. . . . . . . . . . . . . . . . . 32Figure 3.2 XRD spectra of the input slag B. . . . . . . . . . . . . . . . . . . . 33Figure 3.3 Furnace used in pyrometallurgical testing, showing three preloadedcrucibles. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 38Figure 3.4 Crucible used in pyrometallurgical testing. . . . . . . . . . . . . . . 39Figure 3.5 Graphite crucible used for pouring molten samples after pyrometal-lurgical processing. . . . . . . . . . . . . . . . . . . . . . . . . . . . 40Figure 3.6 Leaching setup used. . . . . . . . . . . . . . . . . . . . . . . . . . . 41Figure 3.7 Electrowinning setup, showing the electrodes and HDPE top withslots. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 42Figure 4.1 Lead and silver activity coefficient data and fitted curves [57]. . . . 53Figure 4.2 Tellurium activity coefficient data and fitted curve [58]. . . . . . . . 54xvLIST OF FIGURESFigure 4.3 Ag/Te ratio of the metal phase as a function of temperature. Fixedinputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3. . . 56Figure 4.4 Silver recovery to the metal phase as a function of temperature. Fixedinputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3. . . 57Figure 4.5 Te/Ag ratio of the slag phase as a function of temperature. Fixedinputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3. . . 58Figure 4.6 Tellurium recovery to the slag phase as a function of temperature.Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3. 59Figure 4.7 Ag/Te ratio of the metal phase as a function of temperature and SiO2addition. Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kgof Na2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 60Figure 4.8 Silver recovery to the metal phase as a function of temperature andSiO2 addition. Fixed inputs were 100 kg of slag, 20 kg of PbO, and10 kg of Na2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . . . 61Figure 4.9 Recovery of tellurium to the slag phase as a function of temperatureand SiO2 addition. Fixed inputs were 100 kg of slag, 20 kg of PbO,and 10 kg of Na2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . 62Figure 4.10 Ag/Te ratio of the metal phase as a function of temperature andNa2CO3 addition. Fixed inputs were 100 kg of slag and 20 kg of PbO. 63Figure 4.11 Silver recovery to the metal phase as a function of temperature andNa2CO3 addition. Fixed inputs were 100 kg of slag and 20 kg of PbO. 64Figure 4.12 Tellurium recovery to the slag phase as a function of temperatureand Na2CO3 addition. Fixed inputs were 100 kg of slag and 20 kg ofPbO. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 65Figure 4.13 Ag/Te ratio of the metal phase as a function of temperature and PbOaddition. Fixed inputs were 100 kg of slag and 10 kg of Na2CO3. . . 66Figure 4.14 Silver recovery to the metal phase as a function of temperature andPbO addition. Fixed inputs were 100 kg of slag and 10 kg of Na2CO3. 67Figure 4.15 Tellurium recovery to the slag phase as a function of temperatureand PbO addition. Fixed inputs were 100 kg of slag and 10 kg ofNa2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 68Figure 4.16 Ag/Te ratio of the metal phase as a function of Na2CO3 and PbOaddition. Fixed inputs were 100 kg of slag and a temperature of 1000◦C. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 69xviLIST OF FIGURESFigure 4.17 Silver recovery to the metal phase as a function of Na2CO3 and PbOaddition. Fixed inputs were 100 kg of slag and a temperature of 1000◦C. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 70Figure 4.18 Tellurium recovery to the slag phase as a function of Na2CO3 andPbO addition. Fixed inputs were 100 kg of slag and a temperatureof 1000 ◦C. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 71Figure 4.19 Silver recovery to the metal phase as a function of initial oxygenamount for various temperatures. Fixed inputs were 100 kg of slag,20 kg of PbO, and 10 kg of Na2CO3. . . . . . . . . . . . . . . . . . 73Figure 4.20 Tellurium recovery to the slag phase as a function of initial oxygenamount for various temperatures. Fixed inputs were 100 kg of slag,20 kg of PbO, and 10 kg of Na2CO3. . . . . . . . . . . . . . . . . . 74Figure 4.21 Tellurium percentage in the gas phase as a function of initial oxygenamount for various temperatures. Fixed inputs were 100 kg of slag,20 kg of PbO, and 10 kg of Na2CO3. . . . . . . . . . . . . . . . . . 75Figure 4.22 Oxygen distribution as a function of initial oxygen amount at 1000◦C. Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg ofNa2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 77Figure 4.23 Mass fractions of the four phases examined in the model as functionsof initial oxygen amount at 1000 ◦C. Fixed inputs were 100 kg ofslag, 20 kg of PbO, and 10 kg of Na2CO3. . . . . . . . . . . . . . . 78Figure 4.24 Sodium distribution as a function of initial oxygen amount at 1000◦C. Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg ofNa2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 79Figure 4.25 Silver distribution as a function of initial oxygen amount at 1000 ◦C.Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3. 80Figure 4.26 Tellurium distribution as a function of initial oxygen amount at 1000◦C. Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg ofNa2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 81Figure 4.27 Simulation baseline Ag/Te and Te/Ag ratios. Fixed inputs were 100tonnes/hour of slag, 20 tonnes/hour of PbO, and 10 tonnes/hour ofNa2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 82Figure 4.28 Simulation baseline silver and tellurium recoveries. Fixed inputs were100 tonnes/hour of slag, 20 tonnes/hour of PbO, and 10 tonnes/hourof Na2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 83xviiLIST OF FIGURESFigure 4.29 Deviation from baseline for the Ag/Te ratio of the metal phase asa function of roast and furnace temperature. Fixed inputs were 100tonnes/hour of slag, 20 tonnes/hour of PbO, and 10 tonnes/hour ofNa2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 84Figure 4.30 Deviation from baseline for the silver recovery to the metal phase asa function of roast and furnace temperature. Fixed inputs were 100tonnes/hour of slag, 20 tonnes/hour of PbO, and 10 tonnes/hour ofNa2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 85Figure 4.31 Deviation from baseline for the Te/Ag ratio of the slag phase as afunction of roast and furnace temperature. Fixed inputs were 100tonnes/hour of slag, 20 tonnes/hour of PbO, and 10 tonnes/hour ofNa2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 86Figure 4.32 Deviation from baseline for the tellurium recovery to the slag phaseas a function of roast and furnace temperature. Fixed inputs were100 tonnes/hour of slag, 20 tonnes/hour of PbO, and 10 tonnes/hourof Na2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 87Figure 4.33 Eh-pH diagram of tellurium in solution at 25 ◦C. . . . . . . . . . . 88Figure 4.34 Fractional distribution of Te species as a function of pH. . . . . . . 89Figure 5.1 Silver and tellurium grade in the metal phase as a function of thesilicon to sodium ratio, Q. The temperature was fixed at 1000 ◦Cwith 10 g of PbO addition. Three replicates were conducted for eachtest. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 91Figure 5.2 Silver and tellurium recovery to the metal phase as a function of thesilicon to sodium ratio, Q. The temperature was fixed at 1000 ◦Cwith 10 g of PbO addition. Three replicates were conducted for eachtest. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 92Figure 5.3 Silver to tellurium ratio in the metal phase as a function of the siliconto sodium ratio, Q. The temperature was fixed at 1000 ◦C with 10g of PbO addition. Three replicates were conducted for each test. . 93Figure 5.4 Silver and tellurium recovery to the metal phase as a function of PbOaddition. The temperature was fixed at 1000 ◦C. Three replicateswere conducted for each test. . . . . . . . . . . . . . . . . . . . . . . 94xviiiLIST OF FIGURESFigure 5.5 Tellurium recovery to the slag phase for each center point test. Thefixed inputs were the temperature at 1000 ◦C, 400 ◦C pre-roast for90 minutes, 50 g slag, 10 g PbO addition, and 5 g Na2CO3 addition.Three replicates were conducted per test. . . . . . . . . . . . . . . . 97Figure 5.6 Silver recovery to the metal phase for each center point test. Thefixed inputs were the temperature at 1000 ◦C, 400 ◦C pre-roast for90 minutes, 50 g slag, 10 g PbO addition, and 5 g Na2CO3 addition.Three replicates were conducted per test. . . . . . . . . . . . . . . . 98Figure 5.7 Silver grade of the metal phase for each center point test. The fixedinputs were the temperature at 1000 ◦C, 400 ◦C pre-roast for 90minutes, 50 g slag, 10 g PbO addition, and 5 g Na2CO3 addition.Three replicates were conducted per test. . . . . . . . . . . . . . . . 99Figure 5.8 Total impurity grade of the metal phase for each center point test.The fixed inputs were the temperature at 1000 ◦C, 400 ◦C pre-roastfor 90 minutes, 50 g slag, 10 g PbO addition, and 5 g Na2CO3 addi-tion. Three replicates were conducted per test. . . . . . . . . . . . . 100Figure 5.9 Observed versus predicted tellurium recovery values to the slag phase.The dashed line indicates a perfect correlation. . . . . . . . . . . . . 104Figure 5.10 Observed versus predicted tellurium recovery values to the metalphase. The dashed line indicates a perfect correlation. . . . . . . . . 105Figure 5.11 Predicted tellurium recovery to the metal phase as a function of tem-perature and time. Fixed inputs were 10 g of PbO / 50 g slag, 5 gNa2CO3 / 50 g slag, and a pre-roasting temperature of 400◦C. . . . 106Figure 5.12 Predicted tellurium recovery to the metal phase as a function oftime and Na2CO3 addition. Fixed inputs were 10 g of PbO / 50g slag, 1000 ◦C pyrometallurgical processing temperature, and a pre-roasting temperature of 400 ◦C. . . . . . . . . . . . . . . . . . . . . 107Figure 5.13 Predicted tellurium recovery to the metal phase as a function of timeand roasting temperature. Fixed inputs were 10 g of PbO / 50 g slag,5 g of Na2CO3 / 50 g slag and 1000◦C pyrometallurgical processingtemperature. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 108Figure 5.14 Predicted tellurium recovery to the metal phase as a function of PbOand Na2CO3 additions. Fixed inputs were 1000◦C pyrometallurgicalprocessing temperature for 45 minutes, a pre-roasting temperatureof 400 ◦C. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 109xixLIST OF FIGURESFigure 5.15 Observed versus predicted silver recovery values to the metal phase.The dashed line indicates a perfect correlation. . . . . . . . . . . . . 110Figure 5.16 Predicted silver recovery to the metal phase as a function of temper-ature and PbO addition. Fixed inputs were 45 minutes processingtime, 5 g of Na2CO3 / 50 g slag, and a pre-roasting temperature of400 ◦C. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 111Figure 5.17 Predicted silver recovery to the metal phase as a function of PbO andNa2CO3 additions. Fixed inputs were 1000◦C pyrometallurgical pro-cessing temperature for 45 minutes and a pre-roasting temperatureof 400 ◦C. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 112Figure 5.18 Predicted silver recovery to the metal phase as a function of time androasting temperature. Fixed inputs were 10 g of PbO / 50 g slag, 5g of Na2CO3 / 50 g slag, and a pre-roasting temperature of 400◦C. 113Figure 5.19 Observed versus predicted impurity grade of the metal phase. Thedashed line indicates a perfect correlation. . . . . . . . . . . . . . . 114Figure 5.20 Predicted impurity grade of the metal phase as a function of tempera-ture and Na2CO3 addition. Fixed inputs were 45 minutes processingtime, 10 g of PbO / 50 g slag, and a pre-roasting temperature of 400◦C. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 115Figure 5.21 Predicted impurity grade of the metal phase as a function of time androasting temperature. Fixed inputs were 1000 ◦C pyrometallurgicalprocessing temperature, 10 g of PbO / 50 g slag, and 5 g of Na2CO3/ 50 g slag. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 116Figure 5.22 Predicted impurity grade of the metal phase as a function of PbOand Na2CO3 additions. Fixed inputs were 1000◦C pyrometallurgicalprocessing temperature for 45 minutes and a pre-roasting tempera-ture of 400 ◦C. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 117Figure 5.23 BSE wide view of the metal phase showing dendrite structures. . . . 121Figure 5.24 BSE image and elemental mapping of the Ag-Bi-Pb alloy. Processingconditions were 30 minutes at 1000 ◦C, with additions of 10 g of PbOand 5 g of Na2CO3 per 50 g of slag, and pre-roasting of 90 minutesat 400 ◦C. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 122Figure 5.25 BSE image of the outside surface of a crucible used for pyrometal-lurgical testing. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 123Figure 5.26 BSE image of the slag-crucible interface from a crucible used forpyrometallurgical testing. . . . . . . . . . . . . . . . . . . . . . . . . 124xxLIST OF FIGURESFigure 5.27 Experimental to model comparison showing the silver recovery to themetal phase. The y-axis shows the actual minus predicted results. . 125Figure 5.28 Experimental to model comparison showing the tellurium recoveryto the slag phase. The y-axis shows the actual minus predicted results.127Figure 5.29 XRD comparison of tests performed at temperatures of 900, 1000,and 1100 ◦C. Fixed inputs were a processing time of 45 minutes, 10g of PbO / 50 g slag, 5 g of Na2CO3 / 50 g slag, and a pre-roastingtemperature of 400 ◦C. . . . . . . . . . . . . . . . . . . . . . . . . . 128Figure 5.30 XRD comparison of tests performed for at 15, 45, and 75 minutes.Fixed inputs were 1000 ◦C pyrometallurgical processing temperature,10 g of PbO / 50 g slag, 5 g of Na2CO3 / 50 g slag, and a pre-roastingtemperature of 400 ◦C. . . . . . . . . . . . . . . . . . . . . . . . . . 130Figure 5.31 XRD comparison of tests with 0, 10, and 20 g PbO addition per50 g slag. Fixed inputs were 1000 ◦C pyrometallurgical processingtemperature for 45 minutes, 5 g of Na2CO3 / 50 g slag, and a pre-roasting temperature of 400 ◦C.. . . . . . . . . . . . . . . . . . . . . 131Figure 5.32 XRD comparison of tests with 1, 5, and 9 g Na2CO3 addition per50 g slag. Fixed inputs were 1000 ◦C pyrometallurgical processingtemperature for 45 minutes, 10 g of PbO / 50 g slag, and a pre-roasting temperature of 400 ◦C. . . . . . . . . . . . . . . . . . . . . 133Figure 5.33 Thermodynamic model results showing Ag as Ag2Te as a function oftemperature, PbO, and Na2CO3 additions. . . . . . . . . . . . . . . 135Figure 5.34 Thermodynamic model results showing Te as Na2Te as a function oftemperature, PbO, and Na2CO3 additions. . . . . . . . . . . . . . . 136Figure 5.35 Thermodynamic model results showing Te as PbTe as a function oftemperature, PbO, and Na2CO3 additions. . . . . . . . . . . . . . . 137Figure 5.36 XRD spectrum and species list for the combined leach input. . . . . 138Figure 6.1 Tellurium concentration as a function of slurry solids percentage inpreliminary analysis. . . . . . . . . . . . . . . . . . . . . . . . . . . 141Figure 6.2 Tellurium leach recovery as a function of sodium hydroxide addi-tion, in weight percentage of the input material, for roasted and notroasted runs during preliminary analysis. The solids percentage foreach test was 1% solids in slurry. . . . . . . . . . . . . . . . . . . . . 142Figure 6.3 XRD spectra for the leach input and roasted leach input. . . . . . . 143Figure 6.4 BSE images of a) leach input and b) roasted leach input. . . . . . . 144xxiLIST OF FIGURESFigure 6.5 BSE image of the roasted leach input. . . . . . . . . . . . . . . . . . 145Figure 6.6 Higher magnification BSE image of the roasted leach input, showingthe structure of the sodium tellurite. . . . . . . . . . . . . . . . . . 146Figure 6.7 Solution pH as a function of time for as-received slag and processedslag B leaching. Fixed inputs were the solids percentage at 5%, so-lution temperature of 25 ◦C, with no additives. . . . . . . . . . . . . 147Figure 6.8 Solution pH as a function of time for various slurry solids percentages.Fixed inputs was solution temperature of 25 ◦C with no additives. . 148Figure 6.9 Solution pH as a function of time for initial NaOH concentration.Fixed inputs were the solids percentage at 5% with a solution tem-perature of 25 ◦C. . . . . . . . . . . . . . . . . . . . . . . . . . . . . 149Figure 6.10 Solution pH as a function of time for various leaching solution tem-peratures. Fixed inputs were the solids percentage at 5% with noadditives. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 150Figure 6.11 Tellurium recovery as a function of time for the as-received slags andthe processed slag. Fixed inputs were the solids percentage at 5%,solution temperature of 25 ◦C, with no additives. Two runs wereconducted for each test. . . . . . . . . . . . . . . . . . . . . . . . . . 151Figure 6.12 Tellurium solution concentration after 24 hours for the as-receivedslags and the processed slag. Fixed inputs were the solids percentageat 5%, solution temperature of 25 ◦C, with no additives. Two runswere conducted for each test. . . . . . . . . . . . . . . . . . . . . . . 152Figure 6.13 Tellurium recovery as a function of time for various slurry solidspercentages. Fixed inputs were the solution temperature of 25 ◦C,with no additives. Two runs were conducted for each test. . . . . . 153Figure 6.14 Tellurium solution concentration after 24 hours for various slurrysolids percentages. Fixed inputs were the solution temperature of 25◦C, with no additives. Two runs were conducted for each test. . . . 154Figure 6.15 Tellurium recovery as a function of time for 5 and 10% slurry solids.Fixed inputs were the solution temperature of 25 ◦C, with no addi-tives. Two runs were conducted for each test. . . . . . . . . . . . . 155Figure 6.16 Tellurium solution concentration for 24 and 96 hour leach runs for 5and 10% slurry solids. Fixed inputs were the solution temperatureof 25 ◦C, with no additives. Two runs were conducted for each test. 156xxiiLIST OF FIGURESFigure 6.17 Tellurium recovery as a function of time for various initial concentra-tions of sodium hydroxide. Fixed inputs were the solids percentageat 5% and solution temperature of 25 ◦C. Two runs were conductedfor each test. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 157Figure 6.18 Tellurium solution concentration after 24 hours for various initialconcentrations of sodium hydroxide. Fixed inputs were the solidspercentage at 5% and solution temperature of 25 ◦C. Two runs wereconducted for each test. . . . . . . . . . . . . . . . . . . . . . . . . . 158Figure 6.19 Tellurium recovery as a function of time for various leaching tem-peratures. Fixed inputs were the solids percentage at 5%, with noadditives. Two runs were conducted for each test. . . . . . . . . . . 159Figure 6.20 Tellurium solution concentration after 24 hours for various leachingtemperatures. Each slurry consisted of 5% solids. Fixed inputs werethe solids percentage at 5%, with no additives. Two runs were con-ducted for each test. . . . . . . . . . . . . . . . . . . . . . . . . . . . 160Figure 6.21 Tellurium recovery as a function of time for various initial concen-trations of sodium sulphide. Fixed inputs were the solids percentageat 5% and a solution temperature of 25 ◦C. One run was conductedfor each test. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 161Figure 6.22 Tellurium solution concentration after 24 hours for various initialconcentrations of sodium sulphide. Fixed inputs were the solids per-centage at 5% and a solution temperature of 25 ◦C. One run wasconducted for each test. . . . . . . . . . . . . . . . . . . . . . . . . . 162Figure 6.23 Tellurium recovery as a function of time for roasted and not roastedslags. Fixed inputs were the solids percentage at 5%, solution tem-perature of 25 ◦C, and no additives. One run was conducted foreach roasting test, with two runs conducted for each unroasted slagcounterpart. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 163Figure 6.24 Tellurium solution concentration after 24 hours for roasted and notroasted slags. Fixed inputs were the solids percentage at 5%, solutiontemperature of 25 ◦C, and no additives. One run was conducted foreach roasting test, with two runs conducted for each unroasted slagcounterpart. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 164Figure 6.25 XRD spectra of the leach input and 1% solids leaching residue. . . . 166Figure 6.26 XRD spectra of the leach input and 15% solids leaching residue. . . 166xxiiiLIST OF FIGURESFigure 6.27 XRD spectra of the leach input and 60 g/L NaOH addition leachingresidue. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 167Figure 6.28 XRD spectra of the leach input and 20 g/L Na2S addition leachingresidue. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 168Figure 6.29 BSE images of a) 1x and b) 5x Na2S addition filter papers. . . . . . 170Figure 6.30 Higher magnification BSE image of the 1x Na2S filter paper, showingan agglomeration of residue. . . . . . . . . . . . . . . . . . . . . . . 171Figure 6.31 SE image of residue particles from the 5x Na2S addition cleaning test.172Figure 7.1 Effect of current density and time on tellurium in solution for thepreliminary electrowinning tests. The duration was 6 hours with asolution temperature of 25 ◦C with no additives. . . . . . . . . . . . 174Figure 7.2 Current efficiency as a function of recovery for preliminary electrowin-ning tests. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 176Figure 7.3 Effect of current density on tellurium concentration in solution asa function of time. Solid markers are the experimental data, whilethe dashed lines are the curve fits. Fixed inputs were the solutiontemperature of 25 ◦C and no solution additives. . . . . . . . . . . . 178Figure 7.4 Effect of temperature on tellurium concentration in solution as afunction of time. Solid markers are the experimental data while thedashed lines are the curve fits. Fixed inputs were current density at150 A/m2 and no solution additives. . . . . . . . . . . . . . . . . . . 179Figure 7.5 Effect of Na2S addition for leach solution purification on the telluriumconcentration in solution as a function of time. Solid markers are theexperimental data while the dashed lines are the curve fits. Fixedinputs were current density at 150 A/m2 and a solution temperatureof 25 ◦C. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 180Figure 7.6 Effect of sucrose concentration on tellurium concentration in solutionas a function of time. Solid markers are the experimental data whilethe dashed lines are the curve fits. Fixed inputs were current densityat 150 A/m2 and a solution temperature of 25 ◦C. . . . . . . . . . . 181Figure 7.7 Comparison of sodium sulphide leach solution and regular leach solu-tion on the tellurium concentration in solution as a function of time.Solid markers are the experimental data while the dashed lines arethe curve fits. Fixed inputs were current density at 150 A/m2 and asolution temperature of 25 ◦C. . . . . . . . . . . . . . . . . . . . . . 182xxivLIST OF FIGURESFigure 7.8 Current efficiency as a function of tellurium recovery and currentdensity. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 183Figure 7.9 Current efficiency as a function of tellurium recovery and temperature.184Figure 7.10 Current efficiency as a function of tellurium recovery and sodiumsulphide cleaning. . . . . . . . . . . . . . . . . . . . . . . . . . . . . 185Figure 7.11 Current efficiency as a function of tellurium recovery and sucroseaddition. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 186Figure 7.12 Current efficiency as a function of tellurium recovery and sodiumsulphide leaching. . . . . . . . . . . . . . . . . . . . . . . . . . . . . 187Figure 7.13 Tellurium product grade for each of the 12 electrowinning tests. Thered bar for 150 A/m2 denotes the baseline used to compare tests. Theerror bars show the standard deviation, calculated from the %RSDobtained from the ICP-MS analysis. . . . . . . . . . . . . . . . . . . 189Figure 7.14 Grade of impurity species present in each electrowinning product. . 191Figure 7.15 Impurity grade comparison of standard leach solution and Na2S PLS. 192Figure 7.16 XRD spectra for electrowon tellurium products. . . . . . . . . . . . 193Figure 7.17 InBeam SE image of the tellurium product electrowon at 50 A/m2for 12 hours at room temperature with no solution additives. . . . . 195Figure 7.18 InBeam SE image of the tellurium product electrowon at 150 A/m2for 12 hours at room temperature with no solution additives. . . . . 196Figure 7.19 InBeam SE image of the tellurium product electrowon at 250 A/m2for 12 hours at room temperature with no solution additives. . . . . 197Figure 7.20 InBeam SE image of the tellurium product electrowon at 250 A/m2for 12 hours at room temperature with no solution additives. . . . . 198Figure 7.21 InBeam SE image of the sodium tellurate hydrate impurity in thetellurium product produced at 250 A/m2 for 12 hours at room tem-perature with no solution additives. . . . . . . . . . . . . . . . . . . 199Figure 7.22 InBeam SE image of the tellurium product electrowon at 150 A/m2for 12 hours at 40 ◦C with no solution additives. . . . . . . . . . . . 200Figure 7.23 InBeam SE image of the tellurium product electrowon at 150 A/m2for 12 hours at 50 ◦C with no solution additives. . . . . . . . . . . . 201Figure 7.24 InBeam SE image of the tellurium product electrowon at 150 A/m2for 12 hours at 50 ◦C with no solution additives. . . . . . . . . . . . 202Figure 7.25 InBeam SE image of the tellurium product electrowon at 150 A/m2for 12 hours at 50 ◦C with no solution additives. . . . . . . . . . . . 203xxvLIST OF FIGURESFigure 7.26 InBeam SE image of the tellurium product electrowon at 150 A/m2for 12 hours at room temperature with 0.05 g/L sucrose addition. . 204Figure 7.27 InBeam SE image of the tellurium product electrowon at 150 A/m2for 12 hours at room temperature from the Na2S PLS. . . . . . . . 205Figure 7.28 Scatter plot showing electrowinning grade versus recovery, where thearea of the point relates to the current efficiency. . . . . . . . . . . . 206Figure 7.29 pH of the tellurium solutions as a function of 10 M acid addition.Fixed inputs were 100 mL of leach solution at room temperature. . 207Figure 7.30 pH of the tellurium solutions as a function of normalized acid addi-tion. Fixed inputs were 100 mL of leach solution at room temperature.208Figure 7.31 Percentage of aluminum in solution as a function of 10 M acid addi-tion. Fixed inputs were 100 mL of leach solution at room temperature.209Figure 7.32 Percentage of arsenic in solution as a function of 10 M acid addition.Fixed inputs were 100 mL of leach solution at room temperature. . 210Figure 7.33 Percentage of selenium in solution as a function of 10 M acid addition.Fixed inputs were 100 mL of leach solution at room temperature. . 211Figure 7.34 Percentage of antimony in solution as a function of 10 M acid addi-tion. Fixed inputs were 100 mL of leach solution at room temperature.212Figure 7.35 Percentage of tellurium in solution as a function of 10 M acid addi-tion. Fixed inputs were 100 mL of leach solution at room temperature.213Figure 7.36 Tellurium dioxide product grade and tellurium recovery as a func-tion of normalized acid addition Fixed inputs were 100 mL of leachsolution at room temperature. . . . . . . . . . . . . . . . . . . . . . 214Figure 7.37 BSE image of precipitate produced using 5 mL of 10 M HNO3 into100 mL of Te PLS at room temperature. . . . . . . . . . . . . . . . 215Figure 7.38 BSE image of natrolite surface formed through addition of 5 mL of10 M HNO3 into 100 mL of Te PLS at room temperature. . . . . . 216Figure 7.39 BSE image of additional precipitate produced using 5 mL of 10 MHNO3 into 100 mL of Te PLS at room temperature. . . . . . . . . . 217Figure 7.40 BSE image of precipitate produced using 7.5 mL of 10 M HNO3 into100 mL of Te PLS at room temperature. . . . . . . . . . . . . . . . 218Figure 7.41 BSE image of spherical precipitate produced using 10 mL of 10 MHNO3 into 100 mL of Te PLS at room temperature. . . . . . . . . . 219Figure 7.42 Higher magnification BSE image of precipitates observed in Figure7.41. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 220Figure 7.43 Agglomerated precipitate observed using BSE detector. . . . . . . . 221xxviLIST OF FIGURESFigure 7.44 BSE image agglomerated precipitate surface as seen in Figure 7.43. 222Figure 7.45 BSE image of filter paper after filtration of acidified Te PLS using12.5 mL of HNO3 into 100 mL of leach solution at room temperature. 223Figure 7.46 Higher magnification BSE image of filter paper after filtration. . . . 224Figure 8.1 Flowsheet for the processing of soda slag. . . . . . . . . . . . . . . . 225Figure A.1 The bismuth-lead (Bi-Pb) system [60]. Reprinted with permissionfrom Springer Nature Customer Service Centre GmbH: Springer Jour-nal of Phase Equilibria. The Bi-Pb (Bismuth-Lead) System, N.A. Gokcen. Copyright 1992. . . . . . . . . . . . . . . . . . . . . . . 246Figure A.2 The silver-lead (Ag-Pb) system [61]. Reprinted with permission fromSpringer Nature Customer Service Centre GmbH: Springer Bul-letin of Alloy Phase Diagrams. The Ag-Pb (Silver-Lead) System,Karakaya and Thompson. Copyright 1987. . . . . . . . . . . . . . . 247Figure A.3 The lead-tellurium (Pb-Te) system [62]. Reprinted with permis-sion from Springer Nature Customer Service Centre GmbH: SpringerBulletin of Alloy Phase Diagrams. The Pb-Te (Lead-Tellurium)System, Lin, Hsieh, Sharma, and Chang. Copyright 1989. . . . . . 248Figure A.4 Close up of the PbTe section of Pb-Te phase diagram [62]. Reprintedwith permission from Springer Nature Customer Service Centre GmbH:Springer Bulletin of Alloy Phase Diagrams. The Pb-Te (Lead-Tellurium) System, Lin, Hsieh, Sharma, and Chang. Copyright 1989. 249Figure A.5 The sodium-tellurium (Na-Te) system [63]. Reprinted with permis-sion from Springer Nature Customer Service Centre GmbH: SpringerJournal of Phase Equilibria. The Na-Te (Sodium-Tellurium) Sys-tem, Pelton and Petric. Copyright 1990. . . . . . . . . . . . . . . . 250Figure A.6 The silver-tellurium (Ag-Te) system [64]. Reprinted with permis-sion from Springer Nature Customer Service Centre GmbH: SpringerJournal of Phase Equilibria. The Ag-Te (Silver-Tellurium) Sys-tem, Karakaya and Thompson. Copyright 1991. . . . . . . . . . . . 251Figure A.7 Section of the silver-tellurium (Ag-Te) system [64]. Reprinted withpermission from Springer Nature Customer Service Centre GmbH:Springer Journal of Phase Equilibria. The Ag-Te (Silver-Tellurium)System, Karakaya and Thompson. Copyright 1991. . . . . . . . . . 252xxviiLIST OF FIGURESFigure B.1 Dimensioned drawing of the graphite crucibles used during pyromet-allurgical testing. Dimensions are in millimetres (mm). . . . . . . . 253Figure B.2 Dimensioned drawing showing the electrode used during main serieselectrowinning testing. All dimensions in millimetres. . . . . . . . . 254Figure B.3 Dimensioned drawing showing the electrode holder used during elec-trowinning testing. All units are in millimetres. . . . . . . . . . . . 255Figure C.1 Silver grade of the metal phase as a function of temperature. Fixedinputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3. . . 257Figure C.2 Tellurium grade of the metal phase as a function of temperature.Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.258Figure C.3 Sodium present in silicate species as a function of temperature. Fixedinputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3. . . 259Figure C.4 Tellurium grade of the slag as a function of temperature. Fixedinputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3. . . 260Figure C.5 Te/Ag ratio of the slag phase as a function of temperature and SiO2addition. Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kgof Na2CO3. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 261Figure C.6 Te/Ag ratio of the slag phase as a function of temperature andNa2CO3 addition. Fixed inputs were 100 kg of slag and 20 kg ofPbO. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 262Figure C.7 Te/Ag ratio of the slag phase as a function of temperature and PbOaddition. Fixed inputs were 100 kg of slag and 10 kg of Na2CO3. . . 263Figure C.8 Te/Ag ratio of the slag phase as a function of Na2CO3 and PbOaddition. Fixed inputs were 100 kg of slag and a temperature of1000 ◦C. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 264Figure E.1 Selenium recovery as a function of time for the as-received slags andthe processed slag. Each slurry consisted of 5% solids. . . . . . . . . 268Figure E.2 Selenium solution concentration after 24 hours for the as-receivedslags and the processed slag. Each slurry consisted of 5% solids. . . 269Figure E.3 Selenium recovery as a function of time for various slurry solids per-centages. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 270Figure E.4 Selenium solution concentration after 24 hours for various slurrysolids percentages. . . . . . . . . . . . . . . . . . . . . . . . . . . . . 271Figure E.5 Selenium solution concentration for 24 and 96 hour leach runs for 5and 10% slurry solids. . . . . . . . . . . . . . . . . . . . . . . . . . . 272xxviiiLIST OF FIGURESFigure E.6 Selenium solution concentration after 24 hours for various initial con-centrations of sodium hydroxide. Each slurry consisted of 5% solids. 273Figure E.7 Selenium solution concentration after 24 hours for various leachingtemperatures. Each slurry consisted of 5% solids. . . . . . . . . . . 274Figure E.8 Selenium solution concentration after 24 hours for various initial con-centrations of sodium sulphide. Each slurry consisted of 5% solids. . 275Figure E.9 Selenium solution concentration after 24 hours for roasted and notroasted slags. Each slurry consisted of 5% solids. . . . . . . . . . . . 276Figure E.10 Silicon recovery as a function of time for the as-received slags andthe processed slag. Each slurry consisted of 5% solids. . . . . . . . . 277Figure E.11 Silicon solution concentration after 24 hours for the as-received slagsand the processed slag. Each slurry consisted of 5% solids. . . . . . 278Figure E.12 Silicon recovery as a function of time for various slurry solids per-centages. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 279Figure E.13 Silicon solution concentration after 24 hours for various slurry solidspercentages. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 280Figure E.14 Silicon solution concentration for 24 and 96 hour leach runs for 5 and10% slurry solids. . . . . . . . . . . . . . . . . . . . . . . . . . . . . 281Figure E.15 Silicon solution concentration after 24 hours for various initial con-centrations of sodium hydroxide. Each slurry consisted of 5% solids. 282Figure E.16 Silicon solution concentration after 24 hours for various initial con-centrations of sodium hydroxide. Each slurry consisted of 5% solids. 283Figure E.17 Silicon solution concentration after 24 hours for various leaching tem-peratures. Each slurry consisted of 5% solids. . . . . . . . . . . . . . 284Figure E.18 Silicon solution concentration after 24 hours for various initial con-centrations of sodium sulphide. Each slurry consisted of 5% solids. . 285Figure E.19 Silicon solution concentration after 24 hours for roasted and notroasted slags. Each slurry consisted of 5% solids. . . . . . . . . . . . 286Figure E.20 Aluminum recovery as a function of time for the as-received slagsand the processed slag. Each slurry consisted of 5% solids. . . . . . 287Figure E.21 Aluminum solution concentration after 24 hours for the as-receivedslags and the processed slag. Each slurry consisted of 5% solids. . . 288Figure E.22 Aluminum recovery as a function of time for various slurry solidspercentages. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 289Figure E.23 Aluminum solution concentration after 24 hours for various slurrysolids percentages. . . . . . . . . . . . . . . . . . . . . . . . . . . . . 290xxixLIST OF FIGURESFigure E.24 Aluminum solution concentration for 24 and 96 hour leach runs for5 and 10% slurry solids. . . . . . . . . . . . . . . . . . . . . . . . . . 291Figure E.25 Aluminum solution concentration after 24 hours for various initialconcentrations of sodium hydroxide. Each slurry consisted of 5%solids. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 292Figure E.26 Aluminum solution concentration after 24 hours for various leachingtemperatures. Each slurry consisted of 5% solids. . . . . . . . . . . 293Figure E.27 Aluminum solution concentration after 24 hours for various initialconcentrations of sodium sulphide. Each slurry consisted of 5% solids.294Figure E.28 Aluminum solution concentration after 24 hours for roasted and notroasted slags. Each slurry consisted of 5% solids. . . . . . . . . . . . 295Figure E.29 Arsenic recovery as a function of time for the as-received slags andthe processed slag. Each slurry consisted of 5% solids. . . . . . . . . 296Figure E.30 Arsenic solution concentration after 24 hours for the as-received slagsand the processed slag. Each slurry consisted of 5% solids. . . . . . 297Figure E.31 Arsenic recovery as a function of time for various slurry solids per-centages. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 298Figure E.32 Arsenic solution concentration after 24 hours for various slurry solidspercentages. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 299Figure E.33 Arsenic solution concentration for 24 and 96 hour leach runs for 5and 10% slurry solids. . . . . . . . . . . . . . . . . . . . . . . . . . . 300Figure E.34 Arsenic solution concentration after 24 hours for various initial con-centrations of sodium hydroxide. Each slurry consisted of 5% solids. 301Figure E.35 Arsenic solution concentration after 24 hours for various leachingtemperatures. Each slurry consisted of 5% solids. . . . . . . . . . . 302Figure E.36 Arsenic solution concentration after 24 hours for various initial con-centrations of sodium sulphide. Each slurry consisted of 5% solids. . 303Figure E.37 Arsenic solution concentration after 24 hours for roasted and notroasted slags. Each slurry consisted of 5% solids. . . . . . . . . . . . 304xxxList of AbbreviationsAAS Atomic Absorption SpectroscopyBSE Back-Scattered ElectronsCCD Central Composite DesignCD Current DensityCE Current EfficiencyDAQ Data Acquisition SystemDSC Differential Scanning CalorimetryEDS Energy-Dispersive X-Ray SpectroscopyEDDR Electrodeposition Redox ReplacementEW ElectrowinningHDPE High Density PolyethyleneICP-AES Inductively Coupled Plasma Atomic Emission SpectrometryICP-MS Inductively Coupled Plasma Mass SpectrometrySCE Standard Calomel ElectrodeSHE Standard Hydrogen ElectrodePLS Pregnant Leach SolutionRO Reverse OsmosisSE Secondary ElectronsSEM Scanning Electron MicroscopeTGA Thermogravimetric AnalysisXRD X-Ray DiffractionxxxiList of Symbolsγ Activity CoefficientX Mole fractiona ActivityI Currentt TimeF Faraday’s Constant, 96485 C/molC Coulombsmol Molez ValenceAg/Te Silver-to-Tellurium RatioQ Silicon-to-Sodium Ratio (Si/Na)R RecoveryM Molar (unit), Molar Mass (calculations)m MassT Temperature◦C Degrees Celsius◦F Degrees CelsiusK Degrees KelvinxxxiiList of Chemical FormulaeH HydrogenH2S Hydrogen SulphideH2O WaterH2O2 Hydrogen PeroxideH2SO4 Sulphuric AcidHNO3 Nitric AcidHCl Hydrochloric AcidC CarbonCO Carbon MonoxideCO2 Carbon DioxideO OxygenO2 Oxygen GasOH- Hydroxide IonNa SodiumNaOH Sodium HydroxideNa2S Sodium SuplphideNa2Te Sodium TellurideNa2TeO3 Sodium TelluriteNa2TeO4 Sodium TellurateNa2O Sodium OxideNa2CO3 Sodium Carbonate (Soda Ash)NaHCO3 Sodium Bicarbonate (Baking Soda)xxxiiiList of Chemical FormulaeNaCl Sodium ChlorideNa2Al2Si3O10· 2H2ONatroliteSi SiliconSiO2 Silicon Dioxide (Silica)S SulphurSO2 Sulphur DioxideKCl Potassium ChlorideCa CalciumCaCl2 Calcium ChlorideCu CopperCu2Te Copper TellurideAs ArsenicSe SeleniumAg SilverAg2Te Silver TellurideSb AntimonySb2O3 Antimony TrioxideTe TelluriumTeO2 Tellurium DioxideTeO32- Tellurite IonPb LeadPbTe Lead TelluridePbO Lead(II) OxideBi BismuthxxxivGlossary of NotationActivity The measure of the effective concentration of a species in a mixture undernon-ideal conditions.Assay A measure of the quality and composition of an ore or other material.BarrenSolutionA solution which has undergone an extraction process and as a result hashad most or all of its valuable species removed.Electrowinning An electrochemical process where electricity is used to deposit a givenspecies from solution onto electrodes.Grade The concentration of a species in an ore or other material.Leaching A hydrometallurgical process where a desired species is removed from itshost material via chemical reactions into a liquid.Lixiviant A liquid used in hydrometallurgy, which can selectively remove species fromits host material.Matte A metal sulphide phase formed during high temperature processing of orescontaining sulphur.Normality The measure of the gram or mole equivalents per litre of solution.PregnantLeach SolutionPLS. The solution from a leaching process which contains valuable speciesin solution.Recovery The amount of a product removed from its host material.Slag An oxide rich by-product of high temperature metal processing.xxxvAcknowledgementsI would like to thank my supervisor, Dr. Lukas Bichler, for his encouragement andguidance throughout this research. Thanks to him, I have had the opportunity to broadenmy knowledge into fields in which I previously was unfamiliar with. His knowledge of thematerials field was invaluable when examining results of the experiments, which allowedme to view the data gathered from a different perspective. I would also like to thank Dr.Dimitry Sediako for his support while working under him as a teaching assistant.I would like to thank my parents, Krzysztof and Jadwiga Mackowiak, for their supportand encouraging my studies throughout this degree.Thanks to Teck Metals Ltd. for providing financial support and the materials used forthis project. Special thanks to Mr. Greg Richards, Ms. Rachel Moore, and Ms. HeatherJohnson for all of their support.Thanks to the technicians from the School of Engineering, including Ms. MichelleTofteland, Mr. Durwin Bossy, Mr. Raymond Sedia, and Mr. Praveen Rajan, for theirhelp over the course of my research. Special thanks to Dr. Mark Button, Mr. DavidArkinstall, and Dr. Sudip Shrestha from the UBCO FiLTER Lab for their assistance withSEM microscopy and ICP analysis. Lastly, thanks to Ms. Anita Lam from the UBCVancouver Department of Chemistry for running my XRD samples.Finally, I would like to recognize all of my labmates for their assistance in helpingwith experiments or for bouncing ideas off of during research. Thanks to Tyler Davis,Kyle Lessoway, Anil Prasad, Somi Doja, Levi LaFortune, Justin Mok, Lava Kumar Pillari,Joshua Stroh, Austin Piche, and Alex Gatzke.xxxviChapter 1Introduction1.1 General OverviewThis dissertation outlines the methodology and results obtained for the treatment of atellurium-bearing silver soda slag via a three-step process. Incentive for this research wasto develop a process for the treatment of a metallurgical by-product from a lead smeltingprocess at Trail Operations of Teck Metals Ltd. (Teck). The tellurium and silver containedin the slag present an opportunity for additional value to be extracted from an otherwiseunwanted product. This dissertation will outline the processing of tellurium, followed by areview of pertinent literature relating to the processing of tellurium. Next, results from athermodynamic model and experimental results are shown, along with a flowsheet of theentire process.1.2 TelluriumTellurium (Te) is the 52nd element in the periodic table. It is a brittle, silver-whitemetalloid with an atomic weight of 127.60. Tellurium has six main isotopes which accountfor 99.02% of tellurium. Two of these isotopes, 128Te and 130Te, which make up 65.82% ofTe, are radioactive with extremely long half-lives of 2.2×1024 and 8.2×1020 years, respec-tively. Both of these isotopes decay via double beta decay [1]. The abundance of telluriumon Earth is low, between 1 and 5 parts per billion (ppb). This ranks tellurium 71st inoverall abundance, similar to that of platinum, palladium, and ruthenium [2]. The cosmicabundance of tellurium is higher than that on Earth, where tellurium ranks 40th. Telluriumhad not been detected in stars until 2012, when a group of researchers determined thattellurium is formed in a rapid neutron capture process named r-process [3].1.2.1 UsesIn 2020, the dominant use of tellurium, 40%, was in solar panels as cadmium telluride(CdTe). CdTe panels have the lowest carbon footprint compared to two competing pan-els, at 15.83 gCO2/kWh, compared to CIGS (copper indium gallium selenide) at 21.4411.2. TelluriumgCO2/kWh, and single crystal silicon panels at 27.20 gCO2/kWh [4]. The efficiency ofCdTe panels is comparable to the panels stated previously, at 22.1% compared to 22.9%for CIGS and 27.6% for single crystal Si [5]. The second highest use, at 30%, was in ther-moelectrics, such as infrared sensors, heat pumps, and RTGs (radioisotope thermoelectricgenerators). Metallurgical alloying applications accounted for 15% of consumption, whereTe is added into steel and copper to improve machinability, and as a carbide stabilizer incast irons. The last major application of Te was in rubber vulcanization, which makesup approximately 5% of Te use. The remaining 10% was all classified as ‘other’ such aspigments and photoreceptors [6–8].1.2.2 EconomicsThe historical price of tellurium was obtained from the United States Geological Survey(USGS). Figure 1.1 shows the price of tellurium, adjusted to 2020 US dollars. Figure1.2 shows the world production of tellurium, again obtained from the USGS. Values forproduction after 2003 did not include the US to avoid disclosing proprietary company data[6, 8].Figure 1.1: Price of tellurium in 2020 USD per kg, from USGS [6].21.2. TelluriumFigure 1.2: World production values from USGS, values after 2003 excluded US to avoiddisclosing company data [6, 8].1.2.3 SourcesDue to tellurium’s low abundance in the Earth’s crust, 90% of global Te production isas a by-product of other refining processes, primarily from copper refinery anode slimes.These anode slimes contain between 1 and 4% Te [9]. Tellurium is primarily found in copperporphyry deposits, however two mines, a gold-tellurium mine in China and an epithermaldeposit in Sweden are able to produce Te as a primary resource [6]. The Selenium andTellurium Development Association (STDA) states that 500 tons of copper ore are requiredto produce 1 pound of tellurium [10].1.2.4 Current Processing MethodsProduction of metallic tellurium or other tellurium compounds is done via hydromet-allurgical processing of copper anode slimes. There are three primary methods in whichtellurium can be removed depending on the chemical composition of the slimes. A detailedexplanation of each method can be found in Section 2.3. Each process involves isolationof tellurium either in solution or as a pure compound, which can be isolated and furtherpurified via electrowinning or carbothermic reduction [11].Tellurium is a problematic element in processing operations of other elements due toits ability to form telluride species with several elements including copper, silver, and gold31.3. Research Scope and Objectives[1]. These reactions can have a negative effect on tellurium processing. Consequently,appropriate reagents schemes are necessary to avoid contamination.1.3 Research Scope and ObjectivesThe current research aimed to develop a method for the processing of a tellurium-richsilver soda slag produced by Teck Trail Operations. This slag is currently not processedby Teck, and as such the recovery of valuable elements such as silver is a motivator for theresearch. Due to Teck not currently treating the slag, it was decided that a process exam-ining the complete extraction and separation of silver and tellurium would be examined, asopposed to a detailed examination of a single step, such as pyrometallurgical extraction.This method of looking at the overall process as opposed to a single step would also showareas of the overall process which may need more time to examine and optimize as opposedto others.The slag in examined in the current work is a complex system comprising of at least14 different elements including C, O, Na, Mg, Si, Ca, Cu, As, Se, Ag, Sb, Te, Pb, and Bi.The slag is considered a ‘soda slag’, which means that it contains high contents of sodiumoxide species (such as Na2O, NaHCO3, Na2CO3). While other tellurium materials havebeen examined in the literature, such a slag has not yet been previously examined, and assuch the research is necessary for a better understanding of the processing characteristicsof this slag. The primary objectives of the research were:1) The separation of Ag and Te from the slag.2) Understanding the effects of various processing methods and parameters including py-rometallurgical processing, leaching, and tellurium recovery via electrowinning.3) Development of a process within the constraints of Teck.1.4 Chapter OrganizationThis dissertation contains nine chapters, excluding the introduction.Chapter 2: Literature Review A review of relevant literature relating to the processingand purification of tellurium.Chapter 3: Experimental Methods An outline of the variables, chemicals, and mate-rials is presented. The methodology for each test is discussed.41.4. Chapter OrganizationChapter 4: Thermodynamics A thermodynamic model was constructed for the roast-ing and pyrometallurgical processing of the soda slag. The model inputs and resultsare displayed and discussed.Chapter 5: Results and Discussion - Pyrometallurgy The results from the pyromet-allurgical processing are discussed. Statistical models from the central compositedesign methodology are discussed.Chapter 6: Results and Discussion - Hydrometallurgy The results from the hydromet-allurgical leaching process are discussed. The effect of each input variable is shownon the recovery of various elements.Chapter 7: Results and Discussion - Tellurium Extraction The results from elec-trowinning and tellurium dioxide precipitation are shown. For electrowinning, theeffect of each input variable on the recovery of tellurium is examined.Chapter 8: Results and Discussion - Flowsheet The best results from the previousthree chapters are complied in a flowsheet, outlining the entire process from the slagto the final product. Recommendations for improving the process are also suggested.Chapter 9: Conclusions and Recommended Future Work This chapter summarizesall pertinent findings from the thermodynamics, statistical modelling, and experimen-tal results. Recommendations for future work are also discussed.5Chapter 2Literature Review2.1 OverviewThroughout the duration of the research, a review of pertinent literature related to theextraction of tellurium was conducted on an ongoing basis. This included published liter-ature on the methods of extraction, thermodynamic analysis, or studies on speciation anddistribution of tellurium during processing. The literature was separated into the follow-ing groups: pyrometallurgical processing, hydrometallurgical processing, and extraction oftellurium species.Phase DiagramsPhase diagrams relevant to the materials used in this dissertation can be found inAppendix A and included the Bi-Pb, Ag-Pb, Pb-Te, Na-Te, and Ag-Te systems. Thematerials examined in this research contained high amounts of silver, lead, bismuth, andtellurium. Inclusion of the phase diagrams in the thesis is intended for informative purposes,since it is not possible to completely describe phase evolution in a complex slag using binaryphase diagrams.2.2 Pyrometallurgical ProcessingSodium carbonate (Na2CO3) can be used as a fluxing reagent in pyrometallurgical sys-tems as a method of impurity removal and slag formation. The type of slag used in thecurrent work, i.e. ‘soda slag’, gets its name from the high sodium oxide and/or sodiumcarbonate content. In high temperature systems, sodium carbonate will decompose as seenin the following equation [12]:Na2CO3 → Na2O + CO2 (2.1)Sodium carbonate is considered a basic species due to it donating oxygen atoms to thesolution. Yokokawa (1984) reported that the equilibrium constant for the above reaction62.2. Pyrometallurgical Processingwas 10-5.7 at 1300 K. Thus, under 1 atm of CO2, the activity of Na2O was also 10-5.7. Ifthe activity of Na2O were to drop below this value, then CO2 would be removed from thesolution; however if the activity of Na2O would increase, then CO2 would dissolve into thesolution. Yokokawa carried out a stability examination as a function of slag basicity andoxygen potential, and found that under highly reducing conditions, further decompositionwould occur as the sodium oxide decomposes into gaseous sodium and oxygen gas [12].Kojo et al. examined the thermodynamics of the removal of selenium and telluriumfrom liquid copper via sodium carbonate slag. Samples were prepared from pure reagentsand combined in an alumina crucible. During testing, the oxygen activity was controlled byadding cuprous oxide to the system, or by inserting a graphite rod. Samples were allowedto reach equilibrium before being drawn via alumina tube and quenched. Analysis wasdone using atomic absorption spectroscopy (AAS). The results of their tests can be seen inFigure 2.1. It can be seen that the distribution coefficients of both tellurium and seleniumdecreased to minimum points around an oxygen pressure of 10−7 and 10−6, respectively[13].72.2. Pyrometallurgical ProcessingFigure 2.1: Equilibrium distribution of selenium and tellurium between copper and sodaslag [13]. Reprinted with permission from Springer Nature Customer Service Centre GmbH:Springer Metallurgical and Materials Transactions B. Thermodynamics of removingselenium and tellurium from liquid copper by sodium carbonate slags, Kojo, Taskinen, andLilius. Copyright 1985.The researchers stated that both selenium and tellurium could be removed from themetal into the slag either as cations or anions, and gave the following reactions [13].[Me] + n/2O2(g) = (MeOn) (2.2)[Me] +m/2(Na2O) = (NamMe) +m/4O2(g) (2.3)where Me represents Se or Te, [ ] represents the metal phase, and ( ) the slag phase.Kojo et al. concluded that from a thermodynamic point of view, both selenium andtellurium can be removed from molten copper under reducing conditions via sodium car-bonate slags even at low concentrations. In order to avoid reoxidation of the species, ahighly reducing atmosphere was required. Lastly, they state that the sodium carbonateslag was more effective for the removal of tellurium compared to selenium [13].Luganova-Chnyrenkova examined the effects of certain factors in the removal of minorelements, including tellurium, from Dore´ metal using soda slag. Seven runs were conducted,82.2. Pyrometallurgical Processingwithin each one the Dore´ metal was varied between synthetic and commercial, as well asthe slag composition and oxygen partial pressure. For each test, 80 g of metal and 5 g ofslag were added to magnesia crucibles. Both metal and slag samples were analysed usingatomic adsorption spectroscopy and scanning electron microscopy. Regarding tellurium, itwas found that increasing the partial pressure of oxygen of the gas phase led to an increasein tellurium reporting to the slag. The author also found that at lower amounts of silica,5% as opposed to 10%, the tellurium-slag-gas interaction was increased due to more Na2Opresent to react with Te or TeO2. Further increasing the silica content led to a decrease inthe Na2O activity, as Na2O reacted with the increased amount of silica. In addition to this,an increase in silica may have led to an increase in precious metal retention in the slag dueto the metal entrainment within the slag [14].Alvear et al. performed experiments to determine thermodynamic values for the removalof tellurium and selenium from molten copper using sodium carbonate slag. Experimentalresults were obtained by first producing the samples by mixing pure reagents and heat-ing them at 946 ◦C for one hour. Samples were then analysed via the EMF method.Researchers were able to determine the activity of sodium oxide (Na2O) in Na2O-CO2-(Te/Se)O2 slag under varying CO2 atmospheres, as well as the activities of (Te/Se)O2 inequilibrium with copper. Distribution experiment results showed that the distribution ratioof (Te/Se) changed depending on the oxygen partial pressure. At higher oxygen partialpressures, the following reaction was proposed to occur [15]:2Na(CO3)0.5(l) + (Te/Se)(l) + O2(g) = Na2(Te/Se)O3(l) + CO2(g) (2.4)In comparison, at lower oxygen partial pressures, the tellurium and selenium were ex-pected to form an intermetallic species described by the following reaction [15].2Na(CO3)0.5(l) + (Te/Se)(l) = 2Na2(Te/Se)(l) + CO2(g) + 1/2O2(g) (2.5)Swinbourne et al. examined the distribution of tellurium during the smelting of copperanode slimes. The authors performed both experimental and computational thermody-namic analyses on the system. The input to the experimental testing was a selenide mattecontaining 2% tellurium produced by mixing cathodic silver, copper, selenium, and tel-lurium in a fused silica tube flushed with argon, followed by a slow increase in temperatureto 1100 ◦C over 7 days. Once this temperature was reached, it was held for 3 hours beforethe sample was allowed to air cool to room temperature. The final matte contained 1.95wt.% tellurium. The slag used in the testing was also synthetically produced by mixingsodium carbonate and silica in such a way that the final slag had a SiO2/Na2O ratio of92.2. Pyrometallurgical Processing2. For testing, predetermined amounts of the matte and slag were placed into a conicalrecrystallized alumina crucible and heated to 1100 ◦C. After 30 minutes, a lance was placedinto the melt and oxygen was fed through for predetermined periods of time. The authorsfound that increasing the oxidation of tellurium led to an increase in the tellurium con-centration in the slag. A thermodynamic model was developed using HSC Chemistry 2.0,which aimed to simulate the experimental conditions of their tests. This model was foundto have a good correlation with the experimental results. Using this model, the researchersfound that lowering the smelting temperature led to a decrease in the tellurium reportingto the metal phase. Lastly, it was also seen that the use of air instead of pure oxygen ledto an increase in the tellurium reporting to the metal phase [16].Perez-Tello et al. developed a kinetic model for the oxidation of selenium and telluriumduring industrial production of dore´ silver in a Kaldo furnace. This research was conductedat a operating plant in Nacozari, Mexico. The focus of the research was to determine thekinetics of selenium and tellurium oxidation during the metal refining step of the process.During this step, air injection was done via lances into the Kaldo furnace, with the objectiveof removing selenium as a gaseous oxide. Plant operators reported that during this step,gaseous tellurium dioxide volatilized and contaminated the selenium oxide stream. Moltensamples were taken from the furance every hour and examined via atomic absorption spec-troscopy. Mass balances were also calculated for all elements at the start and end of therefining operation. The researchers found that at the start of the air injection, only seleniumwould volatilize. This volatilization would continue until the Se/Te mass ratio in the bathdropped to 0.55 or lower, at which point tellurium would begin to volatilize. A computedmodel was generated for the system, which was found to be in good agreement with theobserved experimental results. Finally, researchers stated that the current operating con-ditions of the plant were not suitable to the selective separation of selenium and tellurium,stating that the mass ratio at which tellurium volatilization occurred was not dependenton the composition of the metal. They suggested that further studies on the quaternarySe-Te-Ag-O system were necessary [17].Berchenko et al. examined the Pb-Te-O phase diagram and thermal oxidation of leadtellurium for applications in the semiconductor industry. The researchers first developeda ternary phase diagram for the Pb-Te-O system through Gibbs energy calculations, pro-gressively removing pseudobinary species until a final phase diagram was produced, in aneffort to better understand the oxidation of species at higher temperatures. From the phasediagram, it was found that initial oxidation of PbTe to PbTeO3 could occur at room temper-ature. Increasing the temperature to 400 ◦C showed several new phase boundaries develop-ing, but the main oxidation product was still PbTeO3. Further increasing the temperature102.2. Pyrometallurgical Processingto 500 ◦C showed that Pb3TeO5 and Pb5TeO7 developed instead of PbTeO3. Further ox-idation at this temperature led to the formation of Pb2TeO4. Experimental testing usingpure PbTe crystals was then conducted to examine the kinetics of oxide formation in air.In good agreement with the ternary phase diagram, the researchers observed the formationof PbTeO3 at temperatures of 400◦C and below. When held at 400 ◦C over a period of4 hours, researchers observed the growth of both orthorhombic α-PbTeO3 and tetragonalPbTeO3 after 0.5 hours, with the former disappearing at the one hour mark. Continuingto hold over 4 hours led to an increase in the peak intensity of the tetragonal lead telluridephase [18].Kucharski et al. examined the recovery of tellurium from sodium carbonate slag. Theslag consisted of 6.22% tellurium by weight, and upon modification by silica the Te contentwas as low as 4.94%. Four tests, each with varying concentrations of silica addition, wereconducted at 1300 ◦C. The authors noted that at high concentrations of silica, trace amounts(1.23 wt.%) of tellurium were still present in the slag in the form of sodium telluride (Na2Te).This implied that the majority of the tellurium was not reduced to the telluride. This furtherimplied that the sodium oxide (Na2O) produced during sodium carbonate decomposition(Equation 2.1) was strongly bonded in a silicate polyionic species and did not react withtellurium to form sodium telluride. Two proposed reactions were then given, as follows [19]:Na2O(l) + Te(g) + CO = Na2Te(l) + CO2 (2.6)Na2CO3(l) + Te(g) + CO = Na2Te(l) + 2CO2 (2.7)The authors suggested that if reaction 2.6 was dominant, the tellurium content in the slagwould be higher than reported. They stated that if the amount of silica addition increasedsignificantly above the amount of sodium species, then all of the tellurium should be removedfrom the slag, due to the sodium oxide being completely bound to the silicate species. Assuch, they suggested that reaction 2.7 was the main method of sodium telluride formation.Interestingly, as the temperature increases, so too does the amount of tellurium in theslag. This implied that the silicate polyionic species decreased in stability as temperatureincreased [19].Santoso and Taskinen examined liquidus boundary measurements of the Te-O-Na2O-SiO2 system in air. Understanding of this system is of importance to the dore´ smelting ofprecious metals obtained from anode slimes, where tellurium is often present as an impurity.Fluxing agents including sodium carbonate, silica, and sodium borate form a slag into whichthe tellurium reports. Sodium carbonate increases the fluidity of the slag, which in turnensures that metal entrapment in the slag was kept to a minimum. Researchers prepared112.2. Pyrometallurgical Processingsamples by combining high purity SiO2, Na2CO3, and TeO2 into a pellet. This pellet wasthen placed into a vertical tube furnace on a silica plate. Samples were then heated andheld at various temperatures for a fixed duration of time until equilibrium was obtained.The samples were then removed and quenched in an ice water bath. This allowed for thesample to instantly freeze, preserving the various phases present at temperature. Solidphases were noted to be unchanged while liquid phases formed a glassy or microcrystallinephase. These quenched samples were then mounted in a resin and polished, then analysedvia EMPA analysis. The authors found that at silica saturation, the liquidus temperatureof the system was greatly affeted by the sodium oxide concentration. Tellurium oxide wasnot found to have this same effect. Lastly, at fixed SiO2/TeO2 ratios in the liquid, theliquidus temperature was found to decrease from 1200 to 1000 ◦C when additional TeO2 orSiO2 was added to the system [20].Santoso and Taskinen continued research into a similar system. They investigated thephase equilibria of the TeO2-Na2O-SiO2 ternary system in air between 500 and 1200◦C.The reagents, setup, and method were the same as in the previous work conducted by theauthors. From their tests, they were in agreement with previous researchers Troitskii etal. regarding liquidus compositions at TeO2 saturation (i.e. ternary phase boundary cornerregion at 80 wt.%) in the binary TeO2-Na2O. In addition, they suggest that a eutectic mustexist in the binary TeO2-SiO2 system, connected to the TeO2 phase field. Lastly, agreeingwith their previously discussed results, at silica saturation and temperatures between 1000and 1200 ◦C, the liquidus temperature was affected by the Na2O concentration [21].Zhang et al. examined the effects of swelling oxidation roasting and alkaline leachingon the deselenization and detellurization of precious metal ore concentrates. The resultsfrom the roasting are reported in the following text, while the alkaline leaching resultsare reported in Section 2.3. A precious ore metal concentrate containing 10.58 and 57.21%selenium and tellurium, respectively, was used as the input for the research. Swelling agentsNaCl, KCl, CaCl2, Na2CO3, CaCO3, and NaHCO3 were utilized. 40 g of concentrate wasmixed with 0.068 mol of water to ensure total wetting, followed by the addition of theselected swelling agent, in a 1:1 molar quantity. Roasting at 510 ◦C for 2 hours was thenperformed. Analysis of the products was done using ICP, XRF, XRD, and SEM. Theoxidation roasting of the concentrates increased the valence of the elemental selenium andtellurium. After the roasting, researchers examined the gasification of both selenium andtellurium. It was found that no tellurium volatilized during the roast for any of the swellingproducts. Selenium, however, had gasification rates varying between 4.72 and 59.88%, withmost rates being around 40%. The lowest gasification rate, 4.72%, was obtained with aNa2CO3 swelling agent, whereas the highest rate of 59.88% was obtained with CaCO3.122.2. Pyrometallurgical ProcessingUsing no swelling agent resulted in selenium gasification of 19.28%. XRD analysis of theroast products showed that the elemental tellurium had oxidized to tellurium dioxide, whichwas reported to be favourable in the subsequent leaching [22].2.2.1 ThermodynamicsWu et al. examined the microstructure and performed thermodynamic modelling of theAg-Pb-Te system. The researchers prepared 31 ternary alloy compositions using 99.99%pure Ag, Te, and Pb. These 31 compositions were placed into sealed quartz tubes and al-lowed to reach equilibrium at 850 ◦C for 24 hours. Analysis of the alloy was performed usingscanning electron microscopy, wavelength dispersive x-ray spectrometry, x-ray diffraction,and differential scanning calorimetry. The authors stated that there were eight primarysolidification phases, as follows: Ag, γ-Ag2Te, β-Ag2Te, Ag1.9Te, Ag5Te3, PbTe, Te, andPb. At an alloy composition of Ag-4.3at%Pb-62.6%Te, a ternary-eutectic reaction occurredaccording to the following reaction at 337 ◦C [23]:L PbTe + Ag5Te3 + Te (2.8)This eutectic phase was found to be a Ag5Te3 + Te lamellar structure with small (500nm) PbTe, as seen in Figure 2.2. The bright phase was PbTe, embedded in the lamellarstructure of Ag5Te3 (grey) and Te (black).132.3. Hydrometallurgical ProcessingFigure 2.2: The bismuth-lead (Bi-Pb) system [23]. Reprinted with permission from ElsevierElsevier Materials Chemistry and Physics. The microstructure, liquidus projection andthermodynamic modeling of thermoelectric Ag–Pb–Te system, Wu, Foo, Gierlotka, Chen,and Snyder. Copyright 2013.2.2.2 Pyrometallurgical SummaryPyrometallurgical processing of tellurium-bearing materials examined in the followingsection typically revolved around the removal of tellurium from the metal phase, due totellurium being an impurity in copper or dore´ metal. This was achieved through either theoxidation of the tellurium, via oxygen, to form tellurium dioxide, or through the additionof sodium carbonate in order to form sodium telluride or sodium tellurite. These reactionswith sodium were found to be suppressed when substantial quantities of silica were addedto the mixture, due to the preferential reactions between sodium and silica, forming stablespecies.2.3 Hydrometallurgical ProcessingThe leaching of tellurium has been extensively examined as a viable means for theextraction of tellurium. As outlined in Section 1.2.3, the dominant source of telluriumare copper anode slimes. Many authors of the following papers use these slimes as thefeed material. the methods described are either acidic or alkaline in nature, and as suchhave been split into two sections. However, both methods typically produce a solution oftellurium(IV). Tellurium or tellurium species can then be extracted from this solution, as142.3. Hydrometallurgical Processingexamined in Section 2.4.2.3.1 Acidic LeachingHait et al. studied sulphuric acid leaching of anode slime with additives. The anodeslime in question was obtained from the electrorefining plant of the Indian Copper Complex,in Ghatsila, India. The slime contained 3.38% Te. The slime was prepared by size reductionto minus 200 mesh, followed by drying in an air oven to remove moisture. For each leachingtest, 2.5 g of slimes were used with a liquid/solid ratio of 10:1. Preliminary tests using20 vol% acid for 240 minutes at temperatures varying between 30 and 80 ◦C were able torecover only up to 10% of the tellurium. Researchers first examined the effect of manganesedioxide (MnO2) addition, to aid in the oxidation of the tellurium-bearing species accordingto the reaction [24]:Cu2Te + 4Mn→ 2Cu2+ + Te4+ + 4Mn2+ (2.9)With the addition of the MnO2, researchers were able to achieve tellurium recoveries of 66%with 1.25 g of MnO2 and 20 vol% acid at 80◦C for 240 minutes. Fixing the MnO2 additionat 1.25 g and varying the acid from 10 to 30 vol% led to an increase in tellurium recoveryto 75%. Decreasing the temperature from 80 to 30 ◦C, while fixing acid at 20 vol% andMnO2 addition at 1.25 g, led to a decrease in the tellurium recovery from 34.1 to 10.8%.Researchers fit the tellurium data obtained at 80 ◦C and determined that the data fit theash diffusion control model [24]:1− 3(1− x)2/3 + 2(1− x) ∝ t (2.10)where x is a fraction of metals extracted at time t.Next, researchers examined NaCl as a leaching additive. Adding 3 g of NaCl to a 20 vol%acid solution at 80 ◦C led to an increase in tellurium recovery over time, from approximately10% at 30 minutes to 31% at 180 minutes. Finally, when fixing the NaCl addition at 3 gand varying the MnO2 addition, the tellurium recovery was found to increase from 31% at0 g MnO2 to approximately 82% at 2.5 g MnO2. Researchers suggested that the increasedrecoveries may be due to an increase in H+ activity when NaCl was added. Furthermore,the combination of NaCl and MnO2 may have generated Cl2, which is highly oxidizing. Theresearchers gave the following equations as dissolution reactions [24]:MnO2 + 2NaCl + H2SO4 → MnSO4 + Na2SO4 + Cl2 + H2O (2.11)Cu2Te + Cl2 + H2O→ CuCl2 + H2TeO3 + HCl (2.12)152.3. Hydrometallurgical ProcessingRobles-Vega et al. developed an improved hydrometallurgical process for the productionof tellurium from tellurium slag. The researchers pulverized a slag containing 29.94 wt.%tellurium to a size of minus 250 µm. The slag was then leached in distilled water, with 30 gof slag being added per 100 mL of water. This slurry was leached at 50 ◦C for one hour withmechanical agitation. The resulting solution was then filtered and analysed to determinethat it contained 76.20 g/L tellurium, in the form of sodium tellurite. The tellurium leachingefficiency was 85%. Researchers then recovered tellurium using the conventional method,outlined in Equations 2.13 and 2.14. Acid was added to reach a pH of 5.5 to precipitatetellurium dioxide [25]:Na2TeO3(aq) + H2SO4(aq) = Na2SO4(aq) + TeO2(s) + H2O(aq) (2.13)Following this reaction, the resulting tellurium dioxide precipitate was dissolved in an acidsolution containing 2 M H2SO4 and 2 M HCl. Tellurium was in the Te4+ valence state andas such formed aqueous H2TeO3. This solution then had sulphur dioxide (SO2) gas bubbledthrough to reduce the tellurium to its elemental form according to the reaction [25]:H2TeO3(aq) + 2SO2 + H2O(aq) = Te(s) + 2H2SO4(aq) (2.14)This elemental tellurium was assayed and found to have a grade of 97.69% Te. The alter-native method proposed by the researchers instead used 70 mL of 50% hydrogen peroxideper litre of tellurium solution to precipitate sodium tellurate [25]:Na2TeO3(aq) + H2O2(aq) = Na2TeO4(s) + H2O(aq) (2.15)The sodium tellurate precipitate was then dissolved in an H2SO4 / HCl solution, with thesame composition as the conventional method, to produce aqueous H6TeO6. Again, sulhpurdioxide gas was passed through the resulting solution to precipitate elemental tellurium [25]:H6TeO6(aq) + 3SO2 = Te(s) + 3H2SO4(aq) (2.16)This alternative method produced tellurium with a grade of 99.54% Te, an improvementin purity over the conventional method. Authors noted that in both the conventional andalternative method, impurities such as selenium, silicon, and lead were observed to copre-cipitate with the tellurium, suggesting that the SO2 was not selective, and that alternativessuch as electrowinning could be examined [25].Ma and Yang investigated the optimization of microwave-assisted leaching of copperanode slimes using the central composite design (CCD) methodology. A three-factor CCD162.3. Hydrometallurgical Processingwas used, with parameters being mircowave power, ultrasonic power, and time. Slag sampleswere mixed with 150 mL of sulphuric acid and hydrogen peroxide in a three-necked 250 mLflask, keeping a solid / liquid value of 0.12 g/mL. The tellurium bearing species in the slimeswere Cu2Te and Ag2Te, the expected reactions of these two species were as follows [26]:Cu2Te + 4H2O2 + 2H2SO4 = 2CuSO4 + H2TeO3 + 5H2O (2.17)Ag2Te + 3H2O2 + H2SO4 = Ag2SO4 + H2TeO3 + 3H2O (2.18)The researchers found the tellurium leaching efficiency varied from 43.77 to 92.77%. ANOVAwas performed on the results to determine an optimal value for each of the three variablesinvestigated, which would yield the highest leaching efficiency of copper, tellurium, andselenium. Optimal conditions were found to be microwave power of 834 W, ultrasonicpower of 638 W, and a reaction time of 403 seconds. Three verification tests were ran tovalidate the conditions. The leaching efficiencies for Cu, Te, and Se were found to be 98.88±0.56%, 98.96± 0.78%, 98.57± 0.67%, respectively. Researchers stated that the model wasadequate for this system, as the results were close to the predicted values [26].Lu et al. examined the sequential removal of selenium and tellurium from a high-nickel content copper anode slime. The copper slime examined was first crushed to below0.125 mm and analysed via XRD. The tellurium bearing species identified was Cu4SeTe. Apreliminary test was first conducted using 5 g of crushed slag and 0.5 g of sodium carbonate.The species were blended and roasted at 600 ◦C for 2.5 hrs. The resulting product was thenleached using 25 mL of 100 g/L NaOH solution at 75 ◦C for 2.5 hrs. The residue wasthen recovered and leached with 100 mL of 72 g/L sulphuric acid at 80 ◦C for 2 hrs. Theresulting solutions were then assayed, and researchers found that the alkaline leaching steprecovered 78.09% of the Se, while recovering less and 0.011% of the Te. The subsequent acidleaching step saw a tellurium recovery of 79.92%. Further experiments were then conductedon the roasting, alkaline, and acid leaching to optimize parameters. Researchers proposedthe following reactions for the roasting of tellurium-bearing species [27]:Ag2Te + Na2CO3 + 1.5O2 = Na2TeO4 + 2Ag + CO2 ↑ (2.19)Cu2Te + Na2CO3 + 2.5O2 = Na2TeO4 + 2CuO + CO2 ↑ (2.20)XRD analysis was conducted on the solid products or leach residues after each step. In thecalcined alkaline roasting product, the tellurium species was found to be Cu3TeO6, this ledresearchers to suggest Equation 2.21 as a pathway during roasting. Analysis of the alkalineleaching residue found the tellurium species to be CuTeO3, with the proposed reaction172.3. Hydrometallurgical Processingbeing Equation 2.22. Lastly, no tellurium species were observed in the XRD pattern of theacid leach residue. The proposed acid leaching reaction of the tellurium species is seen inEquation 2.23 [27].2Cu4SeTe + 11O2 + 4Ag = 2Ag2SeO4 + 2Cu3TeO6 + 2CuO (2.21)2Cu3TeO6 = 2CuTeO3 + 4CuO + O2 (2.22)2CuTeO3 + 2H2SO4 = 2CuSO4 + 2H2TeO3 (2.23)Shao et al. examined the leaching of tellurium and bismuth using sulphuric acid andferric chloride. The tellurium ore used by the researchers was from the Dashuigou deposit inChina. This ore was assayed and found to contain 1.52% Te and 2.99% Bi. Through XRDanalysis, the tellurium bearing species were found to be daphyllite (Bi2Te2S) and telluricbismuth (BiTe). Preliminary testing was done using solutions containing a fixed 100 g/LFe3+ and varying the H2SO4 concentration between 0 and 250 g/L. The leach time forthese tests was 120 minutes at a temperature of 80 ◦C. Researchers found that increasingthe sulphuric acid concentration to 100 g/L led to a tellurium recovery of 97.83%. Whenfixing the sulphuric acid concentration at 150 g/L, and varying the ferric concentration,it was found that a maximum Te extraction of 97.99% occurred at 60 g/L Fe3+. Furtherincreasing the ferric concentration led to a minor decrease in the tellurium concentration.Researchers gave the following equations as potential reaction pathways for the telluriumspecies found using XRD [28]:Bi2Te2S(s) + 14Fe3+(aq) = 2Bi3+(aq) + 2Te4+(aq) + 14Fe2+(aq) + S(s) (2.24)BiTe(s) + 7Fe3+(aq) = Bi3+(aq) + Te4+(aq) + 7Fe2+(aq) (2.25)Further investigation by the researchers into other variables to determine their effect ontellurium leaching was carried out. Increasing the liquid-to-solid ratio from 4 to 6 mL/gled to an increase in the tellurium extraction from 94% to 98%. Tellurium extraction wasfound to be independent of leaching time. An increase in the leaching temperature from 25to 90 ◦C led to an increase in the tellurium extraction from 92.01 to 99.17% [28].2.3.2 Alkaline LeachingRhee et al. conducted experiments on the leaching and electrowinning of tellurium fromcopper anode slimes. The copper anode slimes, containing 32.9% Cu, 21.7% Te, 26.8% O,and 6.51% S (remainder Se, As, Pb, Fe, Zn, Sb, Mg, Bi) were ground in a ball mill toan average particle size of 2.5 µm. The powder was leached in a 2 L reactor in an NaOH182.3. Hydrometallurgical Processingsolution, with air being bubbled through. The authors gave the following reaction for theleaching of tellurium [29].Te + 2NaOH + O2 = Na2TeO3 + H2O (2.26)Researchers found leaching efficiencies of 98% could be achieved after 3 hours with apulp density of 10 g/L, while for 100 g/L the same efficiency could be obtained after 6hours. Copper concentration in the solution was found to increase with time. The additionof sodium sulphide (Na2S) led to the removal of impurities in the form of sulphides. Theresulting solution was then electrowon, the results of which are discussed in Section 2.4.1[29].Rhee et al. examined the recovery of tellurium from cemented tellurium with the aim ofminimizing waste. Cemented Te was first ground in a ball mill to a mean particle size of 2.5µm, and leached in a NaOH solution with air being bubbled through. Following the leach,the residue was separated and the solution cleaned with 4 g/L Na2S. When examining theeffect of NaOH concentration, researchers found that a solution of 100 g/L NaOH at 80 ◦Cwould extract 99% of the tellurium. When examining the effect of pulp density, researchersfound that leaching efficiencies of over 95% were obtained with pulp densities of 100 to 200g/L, while increasing the pulp density to 400 g/L decreased the Te leaching efficiency toapproximately 80%. Finally, when examining the effect of temperature, researchers foundthat increasing the temperature from 25 to 80 ◦C increased the tellurium extraction from50 to 99%. An interesting effect was observed during the 80 ◦C test, as after 10 minutesthe tellurium extraction dropped, which the researchers attributed to the over-oxidation ofsoluble tellurite (TeO2−3 ) to insoluble tellurate (TeO2−4 ) [30].Fan et al. examined the recovery of tellurium from high-tellurium copper anode slimesvia alkaline pressure leaching. A thermodynamic evaluation alongside experimental resultswas performed. Researchers first constructed a Pourbaix diagram to determine the stabilityof aqueous species. From this, a thermodynamic leaching model using Gibbs free energyminimization was constructed. Only the leaching behaviour of tellurium and selenium wereconsidered. The model assumed gases and aqueous species were ideal, gas pressure was 0.5MPa, equilibrium partial pressure of oxygen was 0.021 MPa, and the temperature was 150◦C. The model showed that both tellurium and selenium dissolved as either tellurate orselenate, respectively. Copper remained in the solid phase. Researchers concluded from themodel that selective leaching of Te and Se from Cu anode slime could be achieved usingalkaline pressure leaching. Experimental testing was conducted using 300 g of raw Cu anodeslimes, with a free sodium concentration of 30 g/L, temperature of 120 ◦C, pressure of 0.5192.3. Hydrometallurgical ProcessingMPa, and a reaction time of 4 hours. It was found that the pressure and temperature hadthe largest impact on the leaching efficiency of Te. Free sodium was not found to have alarge impact. When compared to the model, the researchers stated the results agreed withthe exception of a low leaching efficiency for selenium [31].Guo et al. examined recovering tellurium from high-tellurium materials via alkalinesulphide leaching and sodium sulphite precipitation. The tellurium material for the exper-iments was a by-product of bismuth refining, containing a high concentration of antimony(23.60%), tellurium (11.60%), and lead (11.50%). Other elements present were Na, Bi, Zn,Si, Fe, Al. Leaching experiments were conducted in a 500 mL reactor, and the lixiviant wasprepared by dissolving sodium sulphide in deionized water. Raw material was added to thereactor, where it was stirred at 300 rpm at varying temperatures and times. The telluriumreaction occurring during leaching was proposed to be [32]:Na2TeO4 + 4Na2S + 4H2O = Na2TeS4 + 8NaOH (2.27)Increasing the concentration of Na2S was found to increase the extraction of tellurium.It was found that an increase from 0 to 40 g/L Na2S increased Te extraction from 18 to 89%.Researchers noted that the 18% Te extracted at 0 g/L Na2S was in the water-soluble formof sodium tellurite, Na2TeO3. Tellurium extraction was found to slightly increase (fromapproximately 87 to 90%) as the temperature increased from 30 to 90 ◦C. Lastly, both theliquid/solid ratio and leaching time were found to have little effect on tellurium extraction, inboth cases recovery was at or around 90% and saw only a slight increase, approximately 3%.Researchers then moved on to extraction of tellurium via sodium sulphite precipitation. Thesulphite was added to the leach solution, where tellurium could be precipitated accordingto the following reaction [32].Na2TeS4 + 3Na2SO3 = 3Na2S2O3 + Na2S + Te ↓ (2.28)Zhang et al. examined the effects of swelling oxidation roasting and alkaline leaching onthe deselenization and detellurization of precious metal ore concentrates. The results fromthe swelling oxidation roasting are reported in Section 2.2. The solids from the roasting stepwere leached using a 100 g/L NaOH solution with a liquid to solid ratio of 5:1. Sampleswere leached for 2 hrs at 80 ◦C. Researchers found that the swelling agent NaHCO3 ledto 100% of tellurium leaching, followed by NaCl with 95.98%. The poorest performingagent was CaCO3 at 47.59% leaching efficiency. Due to the effect that the NaCl had onboth the selenium gasification and tellurium leaching, researchers selected it as the swellingagent to perform subsequent tests. Factors such as the roasting temperature, NaCl dosage,202.3. Hydrometallurgical Processingand sulphuric acid dosage were adjusted, and an ideal value for each was then determined.Researchers determined that a NaCl dosage coefficient of 100 (0.1 g NaCl/ 1 g ore) andsulphuric acid dosage of 1.3 (1.3 mg H2SO4 / 1 g ore) led to the highest tellurium leachingrecovery of 97.64% [22].Xu et al. studied the optimization of tellurium and antimony extraction from residuegenerated through alkalkine sulphide leaching of alkaline tellurium slag using Central Com-posite Design (CCD). The residue was first ground to less than 75 µm and dried at 110◦C for 24 hrs. XRD analysis was able to identify the antimony phase as sodium antimonyhydroxide (NaSb(OH)6), however no tellurium phase was able to be identified, which theauthors attributed to a low content of Te compounds and poor crystallization. Using X-rayphotoelectron spectroscopy (XPS), the tellurium species was identified as sodium tellurate(Na2TeO4). A three-factor CCD was developed, varying the Na2S concentration, tempera-ture, and time. The lixiviant used was produced by mixing hydrated sodium sulphide intodeionized water. Solution volume was 200 mL. Upon heating up to the desired temperature,20 g of residue was added, and mechanically mixed using a paddle stirrer. The sodium tel-lurate in the solution would react to form a soluble thiotellurate (TeS2−4 ) according to thepreviously seen Equation 2.27. Thermodynamic analysis of this reaction found a Gibbs freeenergy of -241.40 kJ/mol, implying spontaneity. Through the CCD model, the optimizedleaching values were found to be 175 g/L Na2S at 80◦C for 60 minutes. These condi-tions led to a tellurium recovery of 96.36%. Tellurium was then recovered from solutionusing Na2SO3 reduction according to Equation 2.28. The Gibbs free energy of this reactionwas found to be 24.85 kJ/mol, which the researchers stated implied that certain dynamicconditions were required for it to be carried out [33].2.3.3 Hydrometallurgical SummaryTellurium leaching involving either acid or base methods involve the recovery of telluriuminto solution as either Te(IV) or Te(VI). The tellurite ion, TeO2−3 , was the form typicallyseen in solution, as either tellurous acid (H2TeO3), or sodium tellurite (Na2TeO3), whichis water soluble. Elemental tellurium may be converted into the tellurite form throughleaching with sodium hydroxide. Typically, it was found that lower amounts of solids,higher solution temperatures, and longer leaching times led in the leaching solutions ledto higher recoveries of tellurium. Recovery of the normally insoluble tellurium(VI), seen insodium tellurate (Na2TeO4), could be leached into solution following the addition of sodiumsulphide. The resulting sodium thiotellurate (Na2TeS4) was water soluble and allowed forthe recovery of tellurium into the leaching solution.212.4. Te Recovery2.4 Te Recovery2.4.1 ElectrowinningNapolitano and Mantell examined the preparation of high purity tellurium from analkaline sodium tellurite solution. Tellurium dioxide was dissolved in a solution of sodiumhydroxide, containing between 156 and 188 g of TeO2 per litre. Researchers varied thetemperature, amperage, and different cathode materials. It was observed that the telluriumwas difficult to remove from lead cathodes, and portions of the cathode could contaminatethe tellurium. Iron electrodes showed signs of corrosion, including flakes coming off whenthe tellurium was stripped. Stainless steel was found to be be the ideal cathode, as thetellurium came off easily and without contamination. Researchers found that increasingtemperatures to 115 ◦F (46.1 ◦C) increased the purity of the deposited tellurium to over99.97%. Further increases in temperature caused a decrease in tellurium purity as additionalimpurities were co-deposited. Similarly, increasing the amperage per area caused a decreasein purity. The primary contaminants were tin and lead [34].Handle et al. preformed experimental and statistical analysis on tellurium electrowin-ning. Samples were prepared from tellurium dioxide with intentionally added seleniumand lead impurities. Experiments were conducted with a triple electrode setup, featur-ing one cathode with a deposition surface area of 210 cm2, with electrodes spaced 2.5 cmapart. Variables investigated included the current density, flow rate, tellurium concentra-tion, NaOH concentration, and temperature. Researchers examined the selenium and leadcodeposition and determined which factors had the largest influence. It was found that athigh current densities, a porous, black layer of colloidal tellurium would deposit accordingto reaction 2.29 [35].2Te2− + 3H2O + TeO2−3 = 3Te + 6OH− (2.29)Impurities were reportedly able to be deposited in the tellurium pores. Increased flowrate and Te concentration was found to aid in the development of the porous surface. Highertemperatures were found to accelerate impurity deposition, thus lowering the purity of thetellurium product. Lead was seen to have two opposing effects, which occurred at bothhigh and low current densities. At low current densities and tellurium concentration, leadelectroplated according to reaction 2.30. At higher current densities, the lead reacted withtellurium to form lead telluride (PbTe) according to reaction 2.31, which could becometrapped in the porous Te surface [35].HPbO−2 + H2O + 2e− = Pb + 3OH− (2.30)222.4. Te RecoveryTe2− + HPbO−2 + H2O = PbTe + 3OH− (2.31)Rhee et al. examined the leaching and electrowinning of tellurium from copper anodeslimes. The leaching tests and results are examined in Section 2.3. Researchers foundthat copper deposited before tellurium during electrowinning affected the quality of thefinal product. Tellurium was found to be reduced at -0.5 V (vs. SHE) regardless of Na2Streatment. Removal of copper using Na2S led to the deposition of only tellurium. A highervoltage (4 vs. 3 V) was found to recover tellurium faster, and current efficiencies close to100% could be achieved. The deposition reaction of tellurium was stated to be the following[29].TeO2−3 + 3H2O + 4e− = Te + 6OH E0 = −0.247V (2.32)Broderick et al., in an extension of the work done by his colleague Handle, examined thestrategies for the optimal operation of the tellurium electrowinning process. The methodand materials were similar to that used by Handle et al. in 1996. The examined pa-rameters were the current density, flow rate, Te concentration, NaOH concentration, andtemperature. Numerical modelling based on the CCD allowed the researchers to examinethe effect on the previously mentioned variables on selected response variables. Using themodel results, researchers were able to provide design considerations for optimizing theelectrowinning process. First, low impurity concentrations (<0.125% total Pb + Se) can beachieved by using lower temperatures (<45 ◦C) and low electrolyte concentrations (<193g/L Te), with the downside being high energy consumption (>2500 kWh/ton). To minimizethe downsides of higher current densities, the temperature could be increased, however thisresulted in increased codeposition of Pb with Te. Selenium contamination could be avoidedby using lower current densities; however, as with the previous statement, the codepositionof Pb was seen to increase. Lastly, the removal of Pb prior to electrowinning reduced thecodeposition of Se by 30 to 50% with altered parameters [36].Ha et al. examined the electrowinning of tellurium from alkaline leach liquor. The solu-tion used for testing was a 2.5 M NaOH solution containing 10 mM TeO2. Electrochemicaltests were done using platinum or tellurium disk electrodes, and electrowinning tests weredone using 304 grade stainless steel electrodes. Cathodic polarization and cyclic voltam-metry testing was conducted to determine deposition characteristics. Electrowinning testswere performed to examine the recovery rate at various Te concentrations, voltages, andtemperatures. Growth of tellurium was described as predominantly in the normal direction,followed by needlelike radial growth within the pores of the tellurium. Higher potentials(-0.95 V or less) caused the resulting Te deposit to be structurally unstable due to rapidnormal growth and dissolution to Te2−2 . This weakness caused breakage of the deposited232.4. Te Recoverytellurium due to the solution flow or hydrogen generation at the cathode. It was found thatthe higher voltages allowed for a higher rate of recovery; however this greatly increased thehydrogen generation which caused low current efficiencies. Recoveries of around 50% led tohigher current efficiency of 85-95%, whereas higher recoveries of 90% caused the efficiencyto drop to 50-60%. The researchers recommended a target recovery of 50%, recycling thespent electrolyte and mixing it with fresh solution [37].Montiel-Santilla´n performed electrochemical research on tellurium deposition from anacid sulfate solution. Pure reagents were used as feed materials, and the resulting solutionwas diluted with sulphuric acid until the pH was 5. From past literature, the authors statethat the dissolution of tellurium dioxide in acid conditions forms the stable telluryl ion(HTeO+2 ), this ion was assumed to be the electroactive species. Experiments were conducedin a Pyrex cell with a triple-electrode configuration. Researchers performed several electro-chemical tests including linear sweep voltammetry and potentiodynamic testing. Variablesinvestigated include electrode rotation speed, temperature, and overpotential. Researcherswere able to show that the telluryl reduction process comprises of two steps, a slower adsorp-tion step, reaction 2.33, followed by a reduction step, reaction 2.34. Researchers concludedthat from the kinetic parameters determined, the electochemical process was irreversible at25 ◦C. The reduction kinetics were also found to be improved with increasing tempertaure(any temperature over 25 ◦C) [38].HTeO+2 = (HTeO)+2 )ads (2.33)(HTeO)+2 )ads + 3H+ + 4e− = Te + 2H2O (2.34)Sany examined the optimisation of certain factors on the electrowinning of telluriumusing PLS (partial least square) modelling. Variables considered were the current density,tellurium concentration, and free caustic concentration. Factors such as temperature, agita-tion, and periodic short circuiting (disrupts the diffusion layer to aid in uniform deposition)were kept constant. Pure tellurium dioxide and sodium hydroxide were used to prepare thesolution. Impurities such as Cu, Ag, Se, As, and Sb were also added as oxides to simulatereal conditions. Tests were conducted in 2 L beakers with 1.5 L of electrolyte and stainlesssteel electrodes. Analysis was done via spark source mass spectrometry and glow dischargemass spectroscopy. The author found that the concentration of free caustic soda had thelargest effect on the amount of impurities deposited, behaving positively (increasing concen-tration) for Bi and Ni, and negatively for Ag, Cd, and Pb. The current density was foundto have a negative effect on all impurities, but not to the same extent as the free caustic.The tellurium concentration was not found to be considerably influential on the deposited242.4. Te Recoveryimpurity concentration. The author concluded that a high concentration of free caustic (120g/l), high current densities (350 A/m2) with periodic short circuiting, and high telluriumconcentrations (120 g/l) were ideal conditions for the electrowinning process studied [39].Rudnik and Koz lowski performed electrochemical studies on copper and tellurium co-deposition from an acidic solution. A solution containing 1 mM Cu(NO3)2 and 1 mMTeO2 was prepared with a pH of 0.25. Potentiostatic experiments were conducted with twoplatinum electrodes and an Ag/AgCl reference. Cyclic voltammetry tests were conductedusing platinum and glassy carbon electrodes, again with an Ag/AgCl reference electrode.The researchers found that two copper-tellurium phases deposited based on the potentialapplied. A “low-copper” phase, with Cu/Te ratio of 1.2±0.2, was obtained between poten-tials of -100 to -500 mV, while a “high-copper” phase, Cu/Te ratio of 1.6±0.1 was observedat potentials below -600 mV. The high-Cu deposition was thermodynamically shown to beattributed to the reaction between Te(IV) with Cu(I) ions, while the low-Cu was a reactionof Te(IV) with Cu(II) ions [40].Zhong et al. examined the optimization of the electrowinning process for the productionof 4N tellurium from alkaline solution. The alkaline leach solution used had a Te concen-tration of 64.85 g/L and a NaOH concentration of 2.5 M. The solution was first cleanedthrough the addition of hydrated sodium sulphide. The authors noted that 0.5 g of hydratedNa2S was sufficient to reach 4N Te purity after electrowinning, however Se and Fe contentwas higher than the 4N-Te standard (Nonferrous Industrial Standard YS/T222-2010). In-creasing the Na2S addition to 3 g led to tellurium purity of 99.998%. Further additionsdid not see increases in the purity of the Te. For further removal of impurities such as Seand As, the Te product could be ultrasonically cleaned. The ultrasonication reduced theconcentration of Se and As to less than 2 ppm; however, no other impurity elements wereeffected. Through the use of electrochemical testing, the researchers were able to identify anoptimal current densities of 85 A/m2. Lastly, researchers stated that the following reactiontakes place in the alkaline solution spontaneously, leading to the formation of powdered Tein solution [41]:2Te2−(aq) + 3H2O(l) + TeO2−3(aq) = 3Te(s) + 6OH−(aq) (2.35)Halli et al. examined the electrochemical recovery of tellurium from metallurgical waste.Researchers examined the recovery of tellurium from an acidic solution using both elec-trowinning (EW) and electrodeposition-redox replacement (EDRR). The input leach solu-tion was obtained by treating industrial Dore´ slag with aqua regia, and this solution wasfound to contain 421 ppm Te. Researchers first examined the EDRR method, varying thedeposition time, cut-off time, deposition potential and cut-off potential. It was found that ashort deposition time, 2 seconds, led to the highest recovery of Te, at around 63%. Chang-252.4. Te Recoverying the deposition potential to a higher cathodic potential (-100 mV vs -500 mV) led tohigher recovery of tellurium, and the ideal cut-off potentials for these tests was found to be-150 mV vs SCE. Above cut-off potentials of +200 mV vs. SCE, tellurium was observed toredissolve into solution, whereas at lower cut-off potentials (-50 mV vs. SCE) there was notenough time for tellurium to replace the sacrificial metal. Electrowinning tests were thenconducted over a potential range of -600 to +250 mV. At potentials of less than -245 mV,the tellurium grade of the metal deposit was 12.8% or lower. Increasing the potential to+75 mV led to a tellurium grade of 55.1%. Researchers noted that 200 seconds of EDRRobtained higher purity tellurium than 1200 seconds of EW. This implied that selective Terecovery from the slag PLS could be achieved through optimized EDRR and EW methods.Finally, researchers suggested that at Te solution concentration values above 300 ppm, EWwould be preferred, where EDRR was effective below this threshold. They suggested acombined EW + EDRR method for maximized recovery of tellurium [42].2.4.2 Copper Telluride (Cu2Te) CementationShibasaki et al. examined the recovery of tellurium from a decopperizing leach solution.The solution used in the study was a sulphuric acid solution containing 3.5 kg/m3 (0.35%)Te. This solution would come in contact with copper with the intention of the formation ofcopper telluride according to the reaction [43]:Te(OH)6 + 5Cu + 3H2SO4 = Cu2Te + 3CuSO4 + 6H2O (2.36)A fixed bed reactor was used to test the stripping of copper telluride from a copper plate.The solution was slowly stirred at varying temperature, and the surface conditions of thecopper plate were analysed afterwards along with the amount of copper telluride formed.Researchers found that the copper telluride would easily separate from the plate at tem-peratures of 80 to 90 ◦C, while 60 to 70 ◦C led to difficulty in removing the precipitates.This removal was important as the copper telluride layer did not allow for the solution tocontact fresh copper, halting the cementation reaction. Tests in a laboratory scale fixed bedreactor found that a linear solution velocity of 3.7 cm/s was required to remove the Cu2Teparticles to prevent blockages in the substrate. Lastly, researchers found that the rate oftellurium removal was first order with respect to the tellurium concentration.Wang et al. examined the removal and recovery of tellurium and selenium from a copperslimes leachate. The leachate was produced by the Phelps Dodge Refining Corporation, inEl Paso. The researchers examined the process chemistry after a change in the feedstockto the autoclave caused substantial changes in leaching conditions for the decopperization262.4. Te Recoveryof the solution due to codeposition of selenium. The greater content of selenium and tel-lurium necessitated the use of a significant amount of copper chips and extended batchtime. Researchers first examined the process chemistry, noting that from their observationsselenium could be removed via SO2 reduction, however due to the 80 g/L H2SO4 solutionconcentration, this reduction was not significant for tellurium. Copper telluride formationfrom tellurium(IV) and tellurium(VI) was described as [44]:4Cu + H2TeO3 + 2H2SO4 = Cu2Te + 2CuSO4 + 4H2O (2.37)5Cu + H6TeO6 + 3H2SO4 = Cu2Te + 3CuSO4 + 6H2O (2.38)When examining the kinetics of a laboratory scale test where SO2 was bubbled through thesolution, researchers noted that the selenium reduction occurred rapidly at the start, de-creasing from 2.78 to 0.74 g/L after an hour. Meanwhile, tellurium concentration marginallydecreased from 4.64 to 4.20 g/L. After 5 hours, additional copper chips were added whichaccelerated the tellurium reduction, in the final hour reducing the tellurium concentrationto 0.60 g/L. When examining the effect of temperature, higher temperatures were found tohave a significant impact on the rate of tellurium reduction. A temperature of 92 ◦C ledto a tellurium decrease from 4.63 to 2.13 g/L in 4 hours, whereas at 80 ◦C a reduction ofonly 4.70 to 4.57 g/L Te occurred over the same time period. During this time period, at92 ◦C and higher, a period of fast Te precipitation occurred, which could account for thisdiscrepancy in concentration. Using these results, researchers suggested that a two stageprocess for separately removing selenium and tellurium as copper selenides/tellurides waspossible. Lastly, researchers noted that this two stage process could reduce operation timeand copper consumption [44].Dorfling and du Plessis examined the precipitation of tellurium and other elements fromcopper sulphate leach solution through the addition of thiourea and copper. The inputsolution was a pregnant copper sulphate leach solution obtained from a base metal refinery,prior to the stream entering a selenium and tellurium removal phase. The solution wasfound to contain 33 mg/L of tellurium. Experiments were conducted in a 1 L glass beakersubmerged in a temperature controlled water bath. An overhead stirrer with stainless steelimpeller was utilized to provide agitation. The solution was then heated to the desiredtemperature, upon reaching temperature the other reagents were preheated and introducedto the solution. Tellurium precipitation was found to be independent of temperature. Themaximum tellurium recovery achieved was 12% with 290% excess thiourea at 50 ◦C with4 g/L Cu addition. Higher levels of tellurium precipitation observed at 50 versus 70 ◦Csuggested that the thiourea interacted with the copper, inhibiting the tellurium precipi-272.5. State of Current Literaturetation reactions between Cu and Te. The copper addition was found to have an effecton the tellurium precipitation. At Cu additions of 4 and 8 g/L, Te precipitation rapidlyreached maximum values after 1 and 2 hours, respectively. As the experiment continued,the tellurium precipitation decreased. The authors contributed this observation to two po-tential effects: first, a high supersaturation ratio at the start of testing leading to rapidTe precipitation led to a deficiency of Te in the solution, causing re-dissolution of Te asequilibrium was restored, or second, the precipitated tellurium complex was replaced withanother solution constituent. The authors were unable to confirm either hypothesis. Lastly,the effect of thiourea addition was found to have an effect on tellurium precipitation. At435% excess thiourea, a noticeable amount of tellurium precipitated, whereas none wouldprecipitate in the absence of thiourea, implying that the precipitation of Te6+ by metalliccopper or cuprous ions was not favourable until redution to Te4+ by thiourea occurred [45].2.4.3 Electrowinning SummaryTellurium recovery via electrowinning led to the formation of high purity telluriumproducts under optimal conditions. Stainless steel electrodes were typically used as thetellurium product could be easily removed and would not be contaminated by the electrodematerial. The tellurium deposit itself was reported as being needle-like with several pores,which would lead to impurity depositions within the pores. The electrowinning process couldbe completed using room-temperature solutions, however some researchers saw an increasein the Te purity at slightly elevated temperatures of around 50 ◦C. Higher temperaturesled to the co-deposition of impurity elements, mainly lead. Increasing the current densityto higher values also led to the co-deposition of impurities.2.5 State of Current LiteratureIn the current literature, most of the work was focused on the removal of telluriumfrom either copper anode slimes, or liquid metals such as copper, due to tellurium being animpurity in those systems. A small amount of papers dealt with slag materials similar tothose which will be examined in this dissertation. The slags which were examined typicallyhad lower silver and tellurium contents compared to the current slag to be examined. In thepyrometallurgical papers, the recovery from the slag phase was not often considered, and inmany cases the starting material was copper anode slimes, as previously mentioned. Whilesimilar reactions are likely to occur in the slag to be examined, the literature examined onlyprovides a starting point from which an understanding of the current slag must be builtup. Similar to pyrometallurgical processing, many of the leaching papers dealt with copper282.5. State of Current Literatureanode slimes, and tended to more heavily focus towards acidic leaching. While alkalineleaching papers were found, again the materials were different than that examined here.Many of the papers did not consider the extraction of the insoluble Te(VI) species sodiumtellurate, which may be a large portion of the current slag to be examined. Thus, thecurrent research aims to bridge some of these gaps in knowledge by examining a slag withdifferent composition than that previously seen, and aims to extract two elements from itas effectively as possible.29Chapter 3Experimental MethodsThis chapter outlines the experimental methods, variables, and chemicals used in theresearch.3.1 ChemicalsThe following chemicals were used during various stages of experimentation. All werepurchased from Fisher Scientific, and their names, formulas, CAS Numbers, manufacturer,and Fisher Catalogue ID are provided in in Table 3.1.Table 3.1: Chemicals used during experimentation.Name Formula CAS No. Manufacturer Fisher IDLead(II) Oxide PbO 1317-36-8 Acros Organics AC199340010Silicon Dioxide SiO2 7631-86-9 Fisher S153-3Sodium Hydroxide NaOH 1310-73-2 Fisher S318-5Sodium Sulphide Na2S 1313-82-2 Fisher S426-212Sodium Carbonate Na2CO3 497-19-8 Fisher S263-1Nitric Acid HNO3 7697-37-2 Fisher A200-212Hydrochloric Acid HCl 7647-01-0 Fisher A144-500Sulphuric Acid H2SO4 7664-93-9 Fisher A300-5003.2 SlagsTwo slags of varying composition were provided by Teck Metals for this research. Theslags were assayed at Teck prior to delivery to UBC, with assays provided in Table 3.2. SlagA arrived in February 2017 and it consisted of gravel sized pieces which required furthersize reduction before use. Upon grinding, the powdered slag was light grey in colour. SlagB arrived in November 2017, and arrived pre-crushed, and was darker in colour compared303.2. Slagsto Slag A.Figure 3.1 shows an optical image of the as-received slag B..Table 3.2: Composition (in wt. %) of slags examined in this thesis.Element (wt%) Slag A Slag BAg (oz/ton) 3200 3900Na (%) 14 11Te 31 27C 10.9 26As 0.02 0.23Bi 3.1 5.5Cu 0.99 0.94Pb 1.2 3.6Sb 0.03 0.55Se 0.31 0.2Mg 0.49 0.90Ca 0.08 0.14Si 0.51 1.12313.2. SlagsFigure 3.1: Optical image of the as-received slag B.Figure 3.2 shows the XRD spectra of the input slag B. The primary tellurium specieswere determined to be tellurium (Te), tellurium dioxide (TeO2), and sodium tellurite(Na2TeO3). Other Te species may exist at concentrations below the detection limit of theXRD system, or may be not visible due to other species peak overlapping or if the speciesare amorphous. The primary input species for silver was found to be Ag2Te. In addition tothe XRD analysis conducted for the current research, historical XRD analysis conducted bySGS showed the additional species nahcolite (NaHCO3), rucklidgeite (Bi7.2Pb1.8Te12) andbismuth (Bi).323.3. Variables InvestigatedFigure 3.2: XRD spectra of the input slag B.3.3 Variables InvestigatedThe following sections outline the variables examined for each processing step. Variablesused in both preliminary and main testing are highlighted.3.3.1 Pyrometallurgical VariablesPreliminary TestingPreliminary pyrometallurgical testing was conducted to determine the effect of certainvariables seen in literature or in the thermodynamic model to have an impact on the system.Preliminary pyrometallurgical variables and their values are summarized in Table 3.3 below.333.3. Variables InvestigatedTable 3.3: Preliminary pyrometallurgical testing variables and values.Variable Investigated ValuesSlag Composition A, BSi/Na Ratio As received, 0.1, 0.125, 0.15, 0.175, 0.2PbO Addition 0, 5, 10, 15 gNa2CO3 Addition 0, 5 gTemperature 900, 1000 ◦CTime 30, 45 minsPre-Roast Roast, No RoastPre-Roast Temperature 400 ◦CPre-Roast Time 30, 60, 90, 120 minsMain TestingPyrometallurgical variables examined during the main testing can be found in Table 3.4.These input variables were selected based on results from preliminary testing, as describedin Section 5.1. The maximum temperature tested, 1100 ◦C, was selected based on theresults of the thermodynamic modelling, described in Chapter 4. Constants for each runwere the slag composition used (Slag B), the slag mass (50 g), and the pre-roasting time(90 minutes).Table 3.4: Input variables and values examined during pyrometallurgical tests.Variable Investigated ValuesTemperature 900, 950, 1000, 1050, 1100 ◦CTime 15, 30, 45, 60, 75 minPbO Addition 0, 5, 10, 15, 20 g/50 g slagNa2CO3 Addition 1, 3, 5, 7, 9 g/50 g slagPre-Roast Temperature 300, 350, 400, 450, 500 ◦C343.3. Variables Investigated3.3.2 Hydrometallurgical VariablesPreliminary TestingPreliminary testing was done to examine the effect of certain variables on the leachingprior to planning larger volume main series leach tests. Preliminary leaching variables andvalues are summarized in Table 3.5. The lixiviant in these tests was DI water, with a fixedstirring speed of 300 RPM.Table 3.5: Preliminary leaching variables and values.Variable Investigated ValuesTime 24 hrsTemperature 25 ◦CSlurry Percent Solids by Weight 1, 2, 3, 4, 5, 7.5, 10, 12.5, 15%NaOH Addition 0, 2.5, 12.5, 20 wt.%Roasting None or 200 ◦CMain TestingLeaching input variables examined during the main testing are summarized in Table3.6. These values were selected based on the results of the preliminary testing which can beseen in Section 6.1. A higher slurry solids percentage (30%) was selected to be examinedafter the results of the preliminary tests were discussed with Teck. The effect of time wasexamined during each test via sampling at the stated time. Slurry mass was calculatedprior to the addition of NaOH or Na2S. Constants were the lixiviant (DI water), the slurrymass (500 g), and the stirring rate (300 RPM).353.3. Variables InvestigatedTable 3.6: Input variables and values examined during leaching tests.Variable Investigated ValuesSlurry Solid Concentration 1, 2.5, 5, 10, 15, 30 wt.%Temperature 25, 40, 50, 60 ◦CNaOH Concentration 0, 20, 40, 60, 80, 100 g/LNa2S Concentration 0, 5, 10, 20, 40 g/LPre-Roast Yes or NoTime15, 30, 45 mins1, 2, 4, 6, 12, 24, 48, 72, 96 hrs3.3.3 Electrowinning VariablesPreliminary TestingPreliminary electrowinning variables and values are summarized in Table 3.7. Thesolution temperature was fixed at 25 ◦C for all preliminary tests.Table 3.7: Preliminary electrowinning variables and values.Variable Investigated ValuesTime 2, 4, 6, 8, 10, 12 hrsCurrent Density 50, 150, 250 A/m2Electrode Composition 304 Stainless SteelElectrode Configuration 2 or 3 electrodesElectrode Spacing 3, 5 cmCathodic Surface Area 18, 36 cm2Stirring 0, 60 RPMLeach Solution Composition 2.2, 5.5 g/L TeMain TestingVariables and values used during the main electrowinning tests are summarized in Table3.8. These values were selected based on the preliminary testing results seen in Section 7.1.363.4. Experimental MethodsThe effect of time was examined during each test via sampling at the stated times. Constantsfor all tests were the solution volume (500 mL), electrode spacing (5 cm), cathodic surfacearea (50 cm2), stirring rate (60 RPM), and a leach solution containing 7 g/L Te.Table 3.8: Input variables and values examined during electrowinning tests.Variable Investigated ValuesCurrent Density 50, 150, 250 A/m2Temperature 25, 40, 50 ◦CSucrose Addition 0, 0.05, 0.1 g/LNa2S Pre-Cleaning 1, 2.5, 5x Impurity Concentration (g/L)Na2S Leach Solution Yes or No3.3.4 Tellurium Dioxide PrecipitationAs an alternative method of tellurium recovery, acid could be added to the telluriumleach solution to precipitate a form of tellurium dioxide, α-TeO2, also known as paratellurite.In order to examine the effects of various acids on the precipitate purity, three different acidswere used: hydrochloric, nitric, and sulphuric. The acids were used to lower the solutionpH. Each acid would be diluted to a concentration of 10 M and added in specific quantitiesto the tellurium leach solution. Hydrochloric and nitric acids were added to 100 mL ofleach solution in quantities of 5, 7.5, 10, 12.5, and 15 mL, while sulphuric acid was addedin quantities of 2.5, 5, 7.5, 10, and 15 mL. Differing amounts of sulphuric acid were useddue to differences in the normality of the solutions.3.4 Experimental MethodsA detailed description of the method used for each testing type is provided in thefollowing sections.3.4.1 Pyrometallurgical TestsAll pyrometallurgical tests were conducted in a PSHR223-2118-P kiln. The furnace hada top-loading design with a removable lid to facilitate easy access to the samples, as seenin Figure 3.3. Two K-type thermocouples, one mounted approximately 20 cm above theother, were used to monitor the furnace temperature. A box furnace was used to pre-heat373.4. Experimental Methodsgraphite crucibles for holding molten samples. This box furnace was heated to 180 ◦C inorder to evaporate any moisture present on the crucible.Figure 3.3: Furnace used in pyrometallurgical testing, showing three preloaded crucibles.Samples and reagents for pyrometallurgical testing were weighed out and mechanicallymixed until a homogeneous blend was achieved. The crucibles used were Fisherbrand 100mL glazed porcelain crucibles (Fisher no. SO2225), as seen in Figure 3.4. Crucibles wereplaced into the furnace and pre-heated. Upon reaching the desired temperature, the cru-cibles were removed from the furnace and loaded with samples. In tests with a pre-roastingstep, the blended samples were simply added to the crucibles at room temperature, andallowed to heat up to the desired pre-roasting temperature. Their temperature was mon-itored with a K-type thermocouple placed in the crucible. Upon reaching the roastingtemperature, a timer was started. Upon completion of the timer, the furnace temperaturewas increased to the desired temperature. A thermocouple, inserted into the center of thesample, was used to determine when the samples had reached reaction temperature.383.4. Experimental MethodsFigure 3.4: Crucible used in pyrometallurgical testing.At approximately 2 minutes prior to the end of the test, the pre-heated graphite crucibleswere removed from the box furnace. The crucibles were placed on a concrete brick, anda stainless steel dowel pin was placed at the bottom of the crucible channel to ensure nomaterial would flow out of the crucibles during pouring. A representative graphite crucibleused can be seen in Figure 3.5. A dimensioned drawing of the graphite crucible can be foundin Appendix B, Figure B.1. The loaded crucibles were then removed from the furnace andtheir contents poured into the graphite crucibles. Due to the viscous nature of the slag, ametal scraper bar was used to remove as much slag as possible from the sample crucible.This process continued for the remaining two crucibles.393.4. Experimental MethodsFigure 3.5: Graphite crucible used for pouring molten samples after pyrometallurgical pro-cessing.When all three graphite crucibles were filled with samples, they were allowed to coolto room temperature prior to removal. Samples were typically removed from the graphitecrucibles by pushing the stainless steel pin upwards through the bottom channel, whichforced the sample out. The samples were doubled bagged and stored until further processing.Once the sample crucibles in the furnace had cooled down, they were weighed to determinethe amount of sample which was unable to be recovered.The processed samples were then weighed, and the metal and slag phases were separatedand individually weighed. The slag phase was pulverized with a ring pulverizer to facilitateeasier analysis and as preparation of the subsequent leaching step.Both the metal and slag phases were acid digested and analysed via ICP-OES. Thesilver content of the metal phase was determined via SEM-EDS analysis.3.4.2 Hydrometallurgical TestsLeaching tests were performed in 600 mL Pyrex beakers placed on magnetically stirredhot plates. RO (reverse osmosis) water was added to the beakers in the amount requiredfor each test. For tests examining the effect of a sodium hydroxide solution, the appropriateamount was dissolved into one litre of RO water and this larger batch was used to fill upeach beaker. This larger batch method was used to ensure that the concentration of NaOHin each duplicate test was the same. Weighing the sodium hydroxide was done quicklyin small batches in order to minimize the effects of the hygroscopic nature of the sodium403.4. Experimental Methodshydroxide.When examining the effects of sodium sulphide, in order to avoid the formation of hy-drogen sulphide, the slag was first added to the beaker as it was found that the additionof slag caused the solution pH to reach 12.5 or higher, where H2S does not form. Afterapproximately one minute, the appropriate amount of Na2S was added. If tests were con-ducted at an elevated temperature then the solution was allowed to first heat up on thehotplates, measured using both a digital thermocouple and a glass thermometer which waspermanently inserted into the solution. Lastly, pH probes were secured into the solutionusing an external clamp, and pH data was recorded for the duration of the majority of tests.An image of the leaching setup can be seen in Figure 3.6.Figure 3.6: Leaching setup used.Processed slag for leaching was first prepared by producing a bulk mixed input to ensureas close to homogeneous input material across each test as possible. Pyrometallurigcal testswith similar conditions (i.e., temperature, time, flux addition) were previously pulverizedand a fraction of each sample was combined and mechanically mixed. Three samples of thepulverized bulk input slag were taken for ICP analysis to evaluate homogeneity. Plasticweighboats were used to measure out a specified amount of slag prior to each leaching test.Once all the slag was weighed out, it was poured into the solution in rapid succession, withthe test timer starting when the first slag entered the solution.Sampling of each solution would occur at 15, 30, 45, and 60 minutes, followed by 2, 4, 6,12, and 24 hours. 10 mL of solution were taken per sample, and this was replaced with 10mL of fresh solution appropriate for each test (i.e., RO or NaOH solution). Sampling orderwas conducted in the same order as the slag was added to each solution, in an attempt totake samples at the same time for each trial. The samples taken were filtered through paperfilters to collect any particulate. The resulting solution was sent for ICP-AES analysis.413.4. Experimental MethodsUpon completion of each test, the solutions were filtered overnight using WhatmanGrade 2 paper filters. These filter papers were comprised of cellulose and retained particlesgreater than 8 µm. The resulting solutions were combined to form a bulk mixed PLS inputfor use in electrowinning tests. The remaining residue was collected and analysed via aciddigestion followed by ICP-AES, SEM, and XRD analysis.3.4.3 ElectrowinningElectrowinning tests were carried out using a PARSTAT 4000 Potentiostat/Galvanostat.500 mL of Te PLS was poured into a 600 mL beaker which was then placed onto a magnet-ically stirred hot plate. A 25 mm stirbar was submerged in the Te PLS to provide mixingduring testing, simulating the flow of solution into and out of the electrowinning cell. Animage of the electrowinning setup can be seen in Figure 3.7.Figure 3.7: Electrowinning setup, showing the electrodes and HDPE top with slots.The electrodes used in testing were made of 304 stainless steel, a diagram of which canbe seen in Appendix B, Figure B.2. The electrodes were waterjet cut from 0.81 mm steelsheet, and had their edges sanded down to avoid sharp corners and a rough surface finishafter waterjetting. The faces of the electrodes were polished prior to testing to remove423.4. Experimental Methodsdebris and contaminants.To ensure a constant electrode surface area, 3M 407 electrode tape was used to coverexcess electrode area. The taped electrodes were placed in a high density polyethylene(HDPE) holder, which featured several slots to ensure consistent electrode spacing of 5 cm.A dimensioned drawing can be seen in Appendix B, Figure B.3. Holes in the top of theholder were used to measure solution temperature and to take samples of the solution.The prepared test cell was then attached to the PARSTAT via two leads, one to eachelectrode. Test parameters were controlled using Verastudio software. Samples of thesolution were taken at 2, 4, 6, and 12 hours to examine time-dependent effects.Once the test was complete, the solution was filtered to remove any tellurium which mayhave fallen off of the cathode. The electrodes were kept attached to the HDPE holder, whichwas put into another beaker to allow to dry. Upon drying, the electrodes were weighed.Deposited tellurium was carefully removed from the cathode and placed in small bags. Anytellurium filtered from the solution was then also placed in a separate bag. Samples of thenow barren solution were then taken for further analysis.The barren solution was assayed via ICP-OES, while the deposited and filtered telluriumwere acid digested and assayed via ICP-OES and ICP-MS. The structure of the depositedtellurium was analysed using SEM, and speciation was examined via XRD.Electrowinning CalculationsIn order to determine the current efficiency, Faraday’s Second Law of Electrolysis, specif-ically the constant current form, seen in Equation 3.1 was utilized.m =(I · tF)(Mz)(3.1)where m is the mass of electrodeposited material expected, I current, t time, F Faraday’sconstant (96,485 C/mol), M the molar mass of the element, and z the valence state.The mass obtained from Faraday’s Law could be compared to the actual mass obtainedto determine the current efficiency:CE (%) =mactualmtheoretical· 100% (3.2)3.4.4 Tellurium Dioxide PrecipitationTellurium dioxide precipitation experiments were carried out at room temperature ina fumehood due to the acids used in the testing. First, 50 mL of each acid with 10 Mconcentration were produced by mixing reagent grade acid with RO water. Second, three433.5. Analytical Methods250 mL beakers were filled with 100 mL of the Te PLS. A 25 mm stir bar was used toprovide mixing during testing. Specific volumes of each acid, as previously described inSection 3.3.4, were measured out in graduated cylinders and poured into the Te solution.Next, the combined solutions were allowed to mix for 30 minutes before being filtered viaWhatman Grade 2 filter paper to separate the precipitate from the barren solution. Lastly,the filter papers were dried in an oven at 60 ◦C over 24 hours to aid in the removal ofany residual moisture. Higher temperatures were not used to avoid any potential undesiredreactions. After drying, the filter papers were weighed to determine the mass of the collectedprecipitate.Analysis of the solid precipitate was conducted using SEM-EDS elemental analysis andacid digestion followed by ICP-AES analysis. The barren solution was analysed via ICP-AES.3.5 Analytical MethodsSeveral analytical methods were used to analyse the slag, products, and by-products ofthe process. A brief description of each can be found below.3.5.1 Scanning Electron Microscope (SEM)Imaging of solid samples was conducted using a TESCAN MIRA3 XMU scanning elec-tron microscope. Through the use of a secondary electron (SE), backscattered electron(BSE), or in-beam SE detector, images could be taken at high magnifications and res-olutions. The BSE detector was primarily used for imaging of the slag, processed slag,and metal products as phases in these products could be differentiated by atomic massdifferences. For images taken at higher magnifications (>50,000x), such as those of thefinal tellurium product to see morphological differences, the in-beam SE detector was used.Lastly, an attached Oxford Instruments X-Max EDS detector was used for elemental analy-sis via energy dispersive x-ray spectroscopy (EDS). Point, area, and linescan analyses werepossible using the EDS detector.Samples analysed using the SEM required coating ahead of imaging in order to ensureadequate image quality. The samples, either mounted on carbon stubs or in resin, werecoated with a 10 nm layer of platinum-palladium metal. This layer ensured proper conduc-tivity between the sample and the microscope, ensuring no charging would occur within thesample. When possible, rotation was applied to the samples during the coating process inorder to minimize the effects of any surface imperfections.443.5. Analytical Methods3.5.2 Inductively Coupled Plasma Spectrometry (ICP-AES / ICP-MS)For analysis of both solid and liquid samples, analysis via ICP-AES and ICP-MS wasconducted. Solid sample were first digested in acid using a graphite digestion block. Formost of the samples, 15 mL of a reverse aqua regia solution containing 2:1 HNO3:HCl wasused. For samples where the analysis of silver was necessary, and for the analysis of the puretellurium samples obtained from electrowinning, 20 mL of 1:1 HNO3:DI water solution wasused. In each of the two cases, the solution was digested for two hours at 95 ◦C. Followingdigestion, each sample was diluted to 50 mL before the final dilutions to either 10, 100,or 1000x using either DI water or 1% HNO3 solution. The choice of dilution solution wasdependant on the matrix used in the internal ICP standards and on the sample. Basicsolution samples, either from leaching or electrowinning, were diluted using DI water sincethe addition of small amounts of acid saw the precipitation of tellurium dioxide (TeO2) ornatrolite (Na2Al2Si3O10 · 2H2O) from solution. Samples digested in acid were diluted withthe 1% nitric acid solution.Solution samples from leaching and electrowinning, as well as acid digested solid samplesfrom pyrometallurgical processing and leach residue were analysed using a ThermoFisheriCAP 6000 ICP-AES. This instrument was used for ppm level detection of major elementssuch as Na and Te, and minor elements such as Ag, Se, Bi, Pb, As, Sb, and Cu. Forhigher accuracy in the ppb range, a ThermoFisher X2 quadrupole ICP-MS was used. Thisinstrument was capable of simultaneous analysis of 60 elements at ppt concentration. ICP-MS analysis was done on the electrowon tellurium product to determine product impurityconcentration.3.5.3 X-Ray Diffraction (XRD)X-Ray Diffraction for speciation studies were conducted at the UBC Vancouver De-partment of Chemistry. Powdered samples were sent and the species could be identifiedif they were crystalline in nature. The instrument used was a Bruker D8-Advance X-RayDiffractometer utilizing a Cu radiation source. The sample were packed into a shallow welland were not rotated. The spectra from the XRD analysis was analysed using Qualx2 [46].The databases used were the PDF-2 and Crystallograph Open Database (COD) [47–53].3.5.4 MetallographySamples were prepared for optical and electron microscopy by polishing using abrasivediscs. Samples were first mounted in a resin, where the method of mounting differed if thesamples were solid or powders. Molds 1.25” (3.175 cm) in diameter were used, and solid453.6. Statistical Analysissamples were placed in the bottom of the mold. Then, the resin was poured to fill the mold.Multiple solid samples were also mounted in the same mold using a plastic clip to hold themin place. This clip could allow for up to 5 samples to be mounted in a straight line. Lastly,powder samples could be analysed by first adding the powder to the mold cavity followedby the addition of resin. This resin was then well mixed while still liquid to homogeneouslyblend the powder sample into it. Once this blending occurred, more resin was added tobuild up the height of the sample.Once set, the now mounted samples were removed from the molds and polished usingabrasive silicon carbide (SiC) pads followed by incrementally smaller sized diamond paste.A typical procedure was to polish the samples on 100, 240, 400, and 600 grit SiC paper,then check the surface quality using an optical microscope. Paste with diamond compoundranging from 15 to 3 microns in size was used.Once the samples were polished, they were placed in small bags with their sample IDwritten on them. In the case of hygroscopic samples, silica desiccant was placed in eachbag to avoid unwanted reactions.3.6 Statistical AnalysisTo ensure statistical significance of experimental data between tests, an effort was madeto carry out all tests in triplicate whenever possible. This would minimize the effect ofexperimental bias or random errors in the results. However, due to constraints during sometests only one or two runs were possible. Such tests are indicated when discussing results,and instead of showing variability between tests, variability in the instrument measurementsare shown instead, which acts as a means to show uncertainty in the measurement of a singlesample.3.6.1 Central Composite DesignA Central Composite Design (CCD) methodology was used for developing the exper-iments and analysis of results obtained during pyrometallurgical testing. A five factordesign with 36 total tests, 10 of which were central points (replicates), was generated usingMATLAB®’s ccdesign function. A matrix showing the outputted design, including theorder of experiments conducted, is shown in Table 3.9. Tests with a “C” following the num-ber indicate a center point. Table 3.10 shows the actual variables assigned to each of thevalues seen in the CCD table. Run 24 was performed last, due to the elevated temperatureand greater risk of thermal shock on the crucible.463.6. Statistical AnalysisTable 3.9: Matrix for CCD design generated using MATLAB®, indicating test number.Test numbers with a C are center points.Test X1 X2 X3 X4 X51 C 0 0 0 0 02 -1 -1 -1 -1 13 -1 -1 -1 1 -14 -1 -1 1 -1 -15 C 0 0 0 0 06 -1 -1 1 1 17 -1 1 -1 -1 -18 -1 1 -1 1 19 C 0 0 0 0 010 -1 1 1 -1 111 -1 1 1 1 -112 1 -1 -1 -1 -113 C 0 0 0 0 014 1 -1 -1 1 115 1 -1 1 -1 116 1 -1 1 1 -117 C 0 0 0 0 018 1 1 -1 -1 119 1 1 -1 1 -120 1 1 1 -1 -121 C 0 0 0 0 022 1 1 1 1 123 -2 0 0 0 025 C 0 0 0 0 026 0 -2 0 0 027 0 2 0 0 028 0 0 -2 0 029 C 0 0 0 0 030 0 0 2 0 031 0 0 0 -2 032 0 0 0 2 033 C 0 0 0 0 034 0 0 0 0 -235 0 0 0 0 236 C 0 0 0 0 024 2 0 0 0 0473.6. Statistical AnalysisTable 3.10: Values used in CCD testing.Variable -2 -1 0 1 2X1 - Temperature (◦C) 900 950 1000 1050 1100X2 - Time (min) 15 30 45 60 75X3 - PbO (g/50 g slag) 0 5 10 15 20X4 - Na2CO3 (g/50 g slag) 1 3 5 7 9X5 - Roast Temperature (◦C) 300 350 400 450 5003.6.2 Statistical ModellingStatistical modelling was conducted using MATLAB®, specifically the rstool, fitlmand stepwiselm functions. In these functions, the data was fit to a full quadratic modelaccording to Equation 3.3 [54]. The full quadratic equation contained constant, linear,interaction, and square terms in its output, which provided a method for input variablecomparison.Y = β0 +k∑i=1βiXi +k−1∑i=1k∑j=2(i<j)βijXiXj +k∑i=1βijX2i +  (3.3)The rstools and fitlm functions produced full quadratic equations as outputs; how-ever, these models could be further simplified by removing terms which did not have astatistically significant effect on the models using the stepwiselm function. stepwiselm,as the name suggests, used stepwise linear regression to add or remove terms from the modelbased on a reduction in the sum of squared errors of the model. This was accomplished bydetermining the P-value for the F-test for a model with a term added or removed, based onthreshold values Penter and Premove, which were user-defined and would vary depending onthe system. In most cases, a Penter value of 0.1 and Premove value of 0.25 were used (similarto the values used by Broderick et al. [36]), unless otherwise stated.In order to more easily compare the magnitude of the input variables on the response,the input variables were first normalized by centering and scaling to unit variance, accordingto Equation 3.4.Xsi =Xi −XσX(3.4)Where Xsi is the scaled input value, Xi the input value to be scaled, X the mean of the483.7. Safetyinput data set X, and σX the standard deviation of the input data set X. Scaling the inputdata in such a way removed the effect of the magnitude of the units, such as comparing 1to 9 g of Na2CO3 versus 900 to 1100◦C. Through this scaling, the direct comparison ofcoefficients found by the model was made simpler due to the magnitude of the coefficientreflecting the importance of the coefficient [36].3.7 SafetyDuring every aspect of the research process an emphasis was put on safety. Safety DataSheets (SDS) were kept on file for any material used.As-received slag samples were kept double bagged in pails marked with safety informa-tion and stored on ground level to avoid spillage. Powder or solid samples, such as processedslag or electrowinning products, were kept in marked double bags. Liquid samples such asleach solutions or barren electrowinning solutions were kept in pails which did not reactwith any materials present in the solutions.During pyrometallurgical testing several safety precautions were taken to reduce risks.At all times when handling the as-received slags, labcoats, nitrile gloves, safety glasses,and either a N95 or 3M half-mask respirator was worn. Both furnaces used were situatedin a sand pit which could easily contain any molten spills. A Class D fire extinguisherand fire blanket were present in the area as well. When the furnaces were turned on,any operation with or near them required the use of a heat reflective jacket, reflectiveboot covers, heat resisting gloves, a full face shield, and 3M half-mask respirator. Anypersons entering the casting area were required to wear this gear. While the samples wereprocessed at high temperature, a HEPA vacuum was used to collect any dust generated inthe pyrometallurgical process. During each pouring process, an assistant was present in fullgear as described above to provide any assistant, if needed.Leaching tests lasting over 24 hours had signs present on the lab bench indicating po-tential hazards, emergency contact information, spill clean-up instructions, and emergencyshut-off instructions. Handling of powdered slag samples required a labcoat, gloves, safetyglasses, and a respirator to be worn when handled. Due to the caustic nature of the leachsolutions, handling only occurred when wearing gloves and a labcoat.Electrowinning tests also included signs indicating potential hazards, spill clean-up in-structions, and emergency shut-off instructions. Due to the use of electricity, electricalshock risk signs were also present around the cell. Exposed metal surfaces and connectionpoints were insulated to avoid electric shock.Lastly, when tests were conducted which required the use of sodium sulphide, special493.7. Safetyprecautions were taken to avoid the formation of hydrogen sulphide (H2S) gas. In testscontaining milligram amounts of Na2S, a H2S meter was worn for notification of any con-centration of the gas which may have evolved from the exposure of the Na2S to moisture inthe air. These tests saw the Na2S added to a solution of pH 13, so there was no risk of H2Sformation. During the leaching and electrowinning tests which featured Na2S, a higher riskof H2S formation was present. In order to avoid exposing individuals to the gas, these testswere conducted in a fumehood. Once each test finished the resulting solutions would nolonger pose a risk of H2S generation, but were placed into separate, appropriately labelledcontainers for storage.50Chapter 4ThermodynamicsThis chapter presents results from a thermodynamic model produced in an attemptto predict the outcome of the pyrometallurgical system. An analysis of the behaviour oftellurium in aqueous solution is also presented.4.1 Thermodynamic ConsiderationsA thermodynamic model was developed using the Equilibrium Calculation Module(GEM) of HSC Chemistry 9.1. This module utilizes Gibbs free energy minimization todetermine the equilibrium amounts of selected species at specified conditions. To increasethe accuracy of the model, activity coefficients for select elements published in availableliterature were used [55].The model was used to examine the grade, recovery, distribution, and separation ofimportant elements in the pyrometallurgical processing of the silver soda slag. The effects ofvarious input variables such as temperature, fluxing reagent concentration, and atmosphericcomposition were examined.4.2 Model LimitationsThe thermodynamic model was constructed after preliminary pyrometallurgical testingand product analysis in order to ensure speciation and conditions used in the model wereaccurate. However, there were limitations present in the model. Kinetic analysis was notaccounted for in this model, since the GEM module used only utilizes Gibbs free energy min-imization. As such, any kinetically dependent reactions were not appropriately accountedfor.4.3 Species and Phases ConsideredA total of 66 species were selected from a potential candidate list of 709. Species wereselected due to their stability across the temperature range considered in the model. The514.4. Activity Coefficientsspecies selected were the compounds of the following 14 elements; C, O, Na, Mg, Si, Ca, Cu,As, Se, Ag, Sb, Te, Pb, and Bi. N2 was added to simulate the air atmosphere of the process(partial pressures) and to make calculations easier for the equilibrium solver [55]. The 66selected species were grouped into four phases as follows; Gases, Carbon, Metal, and Slag.Table 4.1 lists the selected species in their respective phases. The phases were selected andpopulated based on the species which were assumed to report to them, where the phaseitself would be a homogeneous blend of the species present. From initial observation of theslag, carbon was observed to be distinct and physically separate from the slag, which waswhy it was segregated to a separate phase in the calculations.Table 4.1: Species and phases considered in the thermodynamic model.Gases Carbon Metal Slag Slag (con’t)CO C Ag Na2Te Na2O2CO2 Na Ag2Te Na2TeO3O2 Te NaTe3 Na2TeO4N2 As Ag2O Na2SeO3SeO2 Bi As2O3 Na2SeO4TeO2 Cu Bi2O3 PbCO3As Pb Bi2Te3 Na2O*2SiO2Te Sb CaCO3 *2Na2O*SiO2Pb Se CaMgSiO4 Na2SiO3PbTe Mg CaO PbONa Ca CaO*MgO Pb3O4Si CaSiO3 PbSiO3PbTe *2CaO*SiO2 SbO2PbBi2Te4 CuO Sb2O3Cu2O Sb2Te3Cu2Te SeO2MgO SiO2Mg2SiO4 TeONa2CO3 TeO2Na2O CaTe4.4 Activity CoefficientsIn order to increase the accuracy of the thermodynamic model, activity coefficients forlead, silver, and tellurium in the silver-lead and silver-tellurium systems were obtained fromliterature. Gierlotka et al. in 2010 examined the Ag-Pb-Te ternary system, utilizing thebinary systems Ag-Te, Pb-Te, and Ag-Pb [56]. Activity coefficients for the slag phase werenot considered in the model. Literature sources that contained activities for slag species524.4. Activity Coefficientswere not used, as the conditions for determining these coefficients, such as temperature ortellurium concentration, were not similar to the current system examined in this disserta-tion. This difference in conditions would result in the extrapolation of the published resultsto work in the model, which may not be accurate.The activity coefficients used in the current model for lead and silver were adapted fromHager and Wilkomirsky. The data provided was plotted as can be seen in Figure 4.1 [57].The MATLAB® curve fitting toolbox was then used to obtain fitted curves. For both leadand silver, rational equations given by Equations 4.1 and 4.2 were obtained.Figure 4.1: Lead and silver activity coefficient data and fitted curves [57].γPb =1.057X2Pb − 0.8151XPb + 0.3434X2Pb − 0.6512XPb + 0.2378· 1200T(4.1)γAg =1.21X2Pb − 1.495XPb + 0.665X2Pb − 1.437XPb + 0.6627· 1200T(4.2)The activity coefficient for tellurium was adapted from Predel and Piehl. Data pointswere plotted as can be seen in Figure 4.2, and the tellurium activity coefficient equation,Equation 4.3, was determined using the MATLAB® curve fitting toolbox [58].534.5. Variables InvestigatedFigure 4.2: Tellurium activity coefficient data and fitted curve [58].γTe =(0.4326X3Ag − 1.072X2Ag − 0.3386XAg + 0.9746)· 1289T(4.3)4.5 Variables InvestigatedThe HSC model examined the effect of the following variables on selected response vari-ables: effect of temperature, silicon dioxide (silica) addition, sodium carbonate addition,lead(II) oxide addition, the effect of a pre-roasting step, and varying the pre-roasting tem-perature. The ranges of the variables were determined through preliminary thermodynamicanalysis, where the terms could be set to arbitrarily large ranges and easily examined. Fromthese results, the ranges were then adjusted to the values examined in the following sections.For models with one variable and the remaining inputs fixed, the middle value from thegiven range was selected (i.e. PbO addition was varied from 0 to 40 kg /100 kg slag so avalue of 20 was used as a fixed value when not examining the effect of PbO). The input slagcomposition used was that of Slag B, discussed in Section 3.2.4.6 Effect of TemperatureThe effect of temperature on the slag system was examined first. In this model, theinputs were fixed as 100 kg of slag, 10 kg of Na2CO3, and 20 kg of PbO. The temperature544.6. Effect of Temperaturewas varied from 800 to 1200 ◦C. The temperature range was selected based on the stability ofthe species examined. At higher temperatures, the thermodynamic data for several specieswould have to be extrapolated for calculation purposes. Both 800 and 1200 ◦C were alsoselected based on the activity coefficients, each of these temperatures were approximately±200 ◦C away from where the coefficients were reported, any temperature higher or lowerthan this would lead to inaccuracies in the model. Additional figures can be found inAppendix C.Metal PhaseFigure 4.3 shows the silver-to-tellurium ratio (Ag/Te) of the metal phase as a func-tion of temperature. The Ag/Te ratio indicates the degree of segregation between silverand tellurium. The objective of the pyrometallurgical process being modelled here was toseparate the silver from the tellurium. Thus, a high Ag/Te ratio implies that the silverwas reporting to the metal phase, while the tellurium reported to the slag. It can be seenfrom Figure 4.3 that as the temperature increased above 900 ◦C, the Ag/Te ratio rapidlyincreased. A maximum ratio of 1503 was obtained at 1016 ◦C. Beyond this temperaturethe ratio dropped, as the species PbTe became more prevalent in the metal phase, and assilver oxidized to form Ag2O in the slag phase.554.6. Effect of TemperatureFigure 4.3: Ag/Te ratio of the metal phase as a function of temperature. Fixed inputs were100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.564.6. Effect of TemperatureFigure 4.4 shows the silver recovery to the metal phase as a function of temperature.The recovery of silver can be seen to increase from 50 to almost 100% between 800 and 975◦C, as the silver oxide or silver telluride species reduce to form elemental silver, which thenreports to the metal phase. At temperatures above 1100 ◦C, this trend reverses, and silverwas seen to report back to the slag phase as Ag2O.Figure 4.4: Silver recovery to the metal phase as a function of temperature. Fixed inputswere 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.574.6. Effect of TemperatureSlag Phase and Tellurium DistributionFigure 4.5 shows the Te/Ag ratio (inverse of Ag/Te ratio) of the slag phase as a functionof temperature. Similar to the Ag/Te ratio, this Te/Ag ratio shows the segregation of eachspecies; however, since the tellurium concentration was higher in the slag, an inverse ratiowas used instead. The peak height was less than half that of the metal ratio, only reaching682 at a slightly higher temperature of 1024 ◦C. The decrease in the ratio above thistemperature was again due to tellurium leaving the slag to form either gaseous Te or intothe metal phase as PbTe, and due to an increase in Ag2O in the slag.Figure 4.5: Te/Ag ratio of the slag phase as a function of temperature. Fixed inputs were100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.584.6. Effect of TemperatureFigure 4.6 shows the recovery of tellurium to the slag phase as a function of temperature.It can be seen that the recovery of tellurium was expected to be above 93% between 800and 1200 ◦C; however, at temperatures between 950 and 1100 ◦C, the recovery was thehighest. At temperatures higher than 1100 ◦C, the decrease in recovery was attributed tothe formation of PbTe species within the metal phase, or to Te volatilization to the gasphase.Figure 4.6: Tellurium recovery to the slag phase as a function of temperature. Fixed inputswere 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.594.7. Effect of Fluxing Reagents4.7 Effect of Fluxing ReagentsThis section examines the effect of each of the four fluxing reagents used in the experi-mental stage of this research and temperature on selected response variables.4.7.1 Effect of Silicon DioxideThe effect of temperature and silicon dioxide addition are examined in this section. Thefixed inputs for this simulation were 100 kg of slag, 10 kg of Na2CO3, and 20 kg of PbO.Metal PhaseFigure 4.7 shows the Ag/Te ratio of the metal phase as a function of temperature andSiO2 addition. It can be seen that increasing the temperature up to 1025◦C leads to alarge increase in the ratio, as previously seen in Figure 4.3. The addition of SiO2, evenin small quantities (i.e., 1 kg per 100 kg slag), had a negative impact on this ratio, sincethe silicon formed sodium silicate species preferentially over the sodium-forming sodium-tellurium species, leading to an increase of Te in the metal phase as PbTe.Figure 4.7: Ag/Te ratio of the metal phase as a function of temperature and SiO2 addition.Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.604.7. Effect of Fluxing ReagentsFigure 4.8 shows the silver recovery to the metal phase as a function of temperatureand SiO2 addition. It can be seen that the addition of SiO2 caused a decrease in thesilver recovery to the metal. This effect can be offset by increasing the temperature of thesystem. This reduction was due to the formation of Ag2Te in the slag phase as a result ofthe preferential formation of sodium silicate species over that of sodium-tellurium species.Figure 4.8: Silver recovery to the metal phase as a function of temperature and SiO2addition. Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.614.7. Effect of Fluxing ReagentsSlag PhaseFigure 4.9 shows the tellurium recovery to the slag phase as a function of temperatureand SiO2 addition. It can be seen that increasing the temperature to between 900 and1100 ◦C with no SiO2 addition yielded the highest recovery of tellurium. The additionof silica, especially at lower temperatures (800-950 ◦C), was seen to lower the recovery oftellurium as the formation of PbTe in the metal phase or gaseous Te occurred (since Tecannot readily form sodium-tellurium species when silica was added to the system due thepreferred formation of sodium silicate species).Figure 4.9: Recovery of tellurium to the slag phase as a function of temperature and SiO2addition. Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.624.7. Effect of Fluxing Reagents4.7.2 Effect of Sodium CarbonateThe effects of temperature and sodium carbonate addition are examined in this section.The fixed inputs for the model were 100 kg of slag and 20 kg of PbO.Metal PhaseFigure 4.10 shows the Ag/Te ratio in the metal phase as a function of temperatureand sodium carbonate addition. It can be seen that an increase in both temperature, andNa2CO3 lead to an increase in this ratio. The addition of sodium carbonate enabled theformation of sodium-tellurium species (such as Na2Te) in the slag phase, while in the processremoving Te from the metal phase.Figure 4.10: Ag/Te ratio of the metal phase as a function of temperature and Na2CO3addition. Fixed inputs were 100 kg of slag and 20 kg of PbO.634.7. Effect of Fluxing ReagentsFigure 4.11 shows the silver recovery to the metal phase as a function of temperature andsodium carbonate. At lower temperatures, it can be seen that the addition of sodium car-bonate had a slight effect on increasing the Ag recovery. This effect was more pronounced athigher temperatures, where the addition of Na2CO3 was necessary to reach higher recoveryvalues (i.e. 4 kg of Na2CO3 at 1050◦C to achieve 95% recovery).Figure 4.11: Silver recovery to the metal phase as a function of temperature and Na2CO3addition. Fixed inputs were 100 kg of slag and 20 kg of PbO.644.7. Effect of Fluxing ReagentsSlag PhaseFigure 4.12 shows the tellurium recovery to the slag phase as a function of temperatureand sodium carbonate addition. Similar to the Ag recovery to the metal phase, it can beseen that an increase in Na2CO3 at lower temperatures led to a slight increase in the Terecovery, with the effect being increased as the temperature increased. Furthermore, theaddition of Na2CO3 counteracted the effect of increasing temperature, thus decreasing theTe recovery above 1000 ◦C.Figure 4.12: Tellurium recovery to the slag phase as a function of temperature and Na2CO3addition. Fixed inputs were 100 kg of slag and 20 kg of PbO.654.7. Effect of Fluxing Reagents4.7.3 Effect of Lead(II) OxideThis section examines the effect of temperature and calcium oxide addition on thesystem. The fixed inputs were 100 kg of slag and 10 kg of Na2CO3.Metal PhaseFigure 4.13 shows the Ag/Te ratio as a function of temperature and lead(II) oxideaddition. It can be seen that the addition of any amount of PbO led to a decrease in theAg/Te ratio, since tellurium was pulled into the metal phase as PbTe.Figure 4.13: Ag/Te ratio of the metal phase as a function of temperature and PbO addition.Fixed inputs were 100 kg of slag and 10 kg of Na2CO3.664.7. Effect of Fluxing ReagentsFigure 4.14 shows the silver recovery to the metal phase as a function of temperatureand lead(II) oxide addition. It can be seen that the addition of any amount of PbO, evenat low temperatures, led to an increase in the Ag recovery due to the conversion of Ag fromthe slag phase species Ag2Te into the metal phase Ag species.Figure 4.14: Silver recovery to the metal phase as a function of temperature and PbOaddition. Fixed inputs were 100 kg of slag and 10 kg of Na2CO3.674.7. Effect of Fluxing ReagentsSlag PhaseFigure 4.15 shows the tellurium recovery to the slag phase as a function of temperatureand lead(II) oxide addition. The addition of PbO led to a decrease in the recovery oftellurium, due to the increased formation of PbTe in the metal phase. This undesirableeffect could be offset by increasing the temperature.Figure 4.15: Tellurium recovery to the slag phase as a function of temperature and PbOaddition. Fixed inputs were 100 kg of slag and 10 kg of Na2CO3.684.7. Effect of Fluxing Reagents4.7.4 Effect of Lead(II) Oxide and Sodium CarbonateThe effect of PbO and Na2CO3 addition on the system is seen in this section. Thetemperature was fixed at 1000 ◦C, with 100 kg of slag input.Metal PhaseFigure 4.16 shows the Ag/Te ratio of the metal phase as a function of sodium carbonateand lead(II) oxide addition. It can be seen that the addition of sodium carbonate led to anincrease in the Ag/Te ratio, while the addition of PbO decreased this ratio. The additionof Na2CO3 led to the formation of sodium-tellurium species (Na2Te and NaTe3) within theslag phase, while the addition of PbO led to the formation of PbTe within the metal phase.Figure 4.16: Ag/Te ratio of the metal phase as a function of Na2CO3 and PbO addition.Fixed inputs were 100 kg of slag and a temperature of 1000 ◦C.694.7. Effect of Fluxing ReagentsFigure 4.17 shows the silver recovery to the metal phase as a function of sodium car-bonate and lead(II) oxide addition. The addition of either of these fluxing reagents ledto an increase in the Ag recovery, however Na2CO3 had a larger effect compared to PbO.This effect was due to the Na2CO3 forming Na2Te and NaTe3 species from the Ag2Te. Theaddition of PbO led to the formation of PbTe from Ag2Te. Both of these interactions ledto the liberation of Ag from Ag2Te, which in the model reported to the metal phase.Figure 4.17: Silver recovery to the metal phase as a function of Na2CO3 and PbO addition.Fixed inputs were 100 kg of slag and a temperature of 1000 ◦C.704.7. Effect of Fluxing ReagentsSlag PhaseFigure 4.18 shows the tellurium recovery to the slag phase as a function of sodiumcarbonate and lead(II) oxide addition. The addition of Na2CO3 led to an increase in therecovery, due to the formation of sodium-tellurium species within the slag. The addition ofPbO led to a higher amounts of PbTe formation within the metal phase, removing telluriumfrom the slag phase.Figure 4.18: Tellurium recovery to the slag phase as a function of Na2CO3 and PbO addition.Fixed inputs were 100 kg of slag and a temperature of 1000 ◦C.714.8. Effect of Initial Oxygen4.8 Effect of Initial OxygenThis section examines the effect of the initial oxygen amount on the system. The initialoxygen amount was important to examine as the system may perform better under differentatmospheres, such as inert N2, or with an oxygen-enriched atmosphere. Figure 4.19 showsthe silver recovery to the metal phase as a function of initial oxygen concentration for variouspyrometallurgical processing temperatures. It can be seen that for temperatures of 900 ◦Cand higher, the Ag recovery was at or close to 100% until 0.7 kmol of oxygen were added.After this point, the Ag recovery decreased as the oxygen amount increased. The recoveryfor all temperatures reaches 0% at a value of approximately 2.1 kmol input O2, after whichthe Ag recovery sharply increased. The exact mechanisms behind the changes are explainedfurther in Section 4.8.1; however, in summary the first decrease of Ag recovery was due tothe formation of Ag2Te, whereas the later increase was due to the oxidation of the Te inthe Ag2Te to form TeO, TeO2, and Ag metal.724.8. Effect of Initial OxygenFigure 4.19: Silver recovery to the metal phase as a function of initial oxygen amount forvarious temperatures. Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.734.8. Effect of Initial OxygenFigure 4.20 shows the tellurium recovery to the slag phase as a function of the initialoxygen amount for various temperatures. It can be seen that for each temperature, theTe recovery to the slag phase decreased as the initial oxygen amount increased, reaching aminimum value at 1.9 kmol input O2. The recovery increased when more O2 was added.The decrease in Te recovery at all temperatures was due to the formation of other telluriumspecies, either gaseous ones at higher temperatures, as seen in Figure 4.21, showing thepercentage of tellurium as a gas as a function of input oxygen for various temperatures.Alternatively, the formation of elemental tellurium or PbTe could occur, both of whichwere assumed to report to the metal phase. Once a sufficient amount of oxygen was presentin the system, and with nothing else to react with (as explained later in this section) theformation of tellurium oxides occurred, which forced Te back into the slag phase.Figure 4.20: Tellurium recovery to the slag phase as a function of initial oxygen amountfor various temperatures. Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg ofNa2CO3.744.8. Effect of Initial OxygenEvolution of gaseous Te, in Figure 4.21, was seen to follow an inverse trend of the slagTe recovery. The amount of Te in the gas phase increased as the Te recovery to the slagphase decreased, which was expected due to the Te vapourizing. This volatilization wasincreased at higher temperatures (such as 1200 ◦C), where the majority of the tellurium notin the slag phase was reporting to the gaseous phase. The difference between the gas andthe slag recovery was Te in the form of PbTe or elemental Te, both of which were assumedto report to the metal phase.Figure 4.21: Tellurium percentage in the gas phase as a function of initial oxygen amountfor various temperatures. Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg ofNa2CO3.754.8. Effect of Initial Oxygen4.8.1 Element Distribution at 1000 ◦CFigures 4.22 through 4.26 examine the effect of the initial concentration of oxygen onthe speciation of different elements. Only species with the highest distributions are shown,therefore minor species are not included which may result in the distributional sum notbeing equal to 100%. In addition, two vertical lines are present on each plot. The first linemarked at 0.74 kmol O2 input represents the point at which all free carbon was removedfrom the system, and was being replaced by carbon monoxide. To the left of this line, therewas still free carbon present; however, it diminished as more oxygen was added. The secondline marked at 1.91 kmol O2 input represents the point at which the carbon monoxide hadfully converted to carbon dioxide.Figure 4.22 shows the oxygen distribution between three major species as a function ofinitial oxygen amount at 1000 ◦C. It can be seen that the CO percentage increased untilthe first vertical line at 0.74 kmol O2 input, during this phase there was free carbon presentin the system, which was being converted to CO as the input O2 increased, according tothe equation:2C + O2 = 2CO (4.4)Once all the free carbon was removed, the addition of further oxygen began converting theCO into CO2, and produced Na2CO3, albeit to a lesser extent. The conversion from COto CO2 continued until all the CO was converted, at 1.91 kmol O2 input, according to theequation:2CO + O2 = 2CO2 (4.5)After this point, the oxygen could no longer form CO2, forcing it to react with otherspecies to form various oxides, which are examined in the following figures. In Figure 4.22,the continued addition of oxygen can be seen to have resulted in a linear decrease in thepercentage distribution of oxygen to CO2, as species were converted to their oxide forms.764.8. Effect of Initial OxygenFigure 4.22: Oxygen distribution as a function of initial oxygen amount at 1000 ◦C. Fixedinputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.774.8. Effect of Initial OxygenFigure 4.23 shows the mass fraction of each phase as a function of initial oxygen amountat 1000 ◦C. It can be seen that the carbon phase decreased to 0% by the first vertical line at0.74 kmol O2 input. Once this occurred, the slag phase fraction increased, while the metalphase fraction decreased, due to the oxidation of the metal species, which then reported tothe slag phase. As the initial oxygen amount increased, the CO fully converted to CO2.After this point, oxidation of the metal phase occurred, as seen by the decrease in the metalphase distribution to 0% while the slag phase increased to a maximum of 58%. Once thisminimum value of metal fraction was reached, the mass fraction of metal then increased, asthe metal phase reformed due to the conversion of Ag2Te back to Ag, as explained furtherin Figure 4.25.Figure 4.23: Mass fractions of the four phases examined in the model as functions of initialoxygen amount at 1000 ◦C. Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg ofNa2CO3.784.8. Effect of Initial OxygenFigure 4.24 shows the distribution of sodium as a function of initial oxygen amount at1000 ◦C. It can be seen that before the first vertical line, there was little change in thesodium distribution besides that of the gas and carbonate phase, which behaved inverselyof each other. In the gas phase, gaseous sodium (Na(g)) was removed as more CO was gen-erated, forming sodium carbonate (Na2CO3). Once no free carbon remained, the formationof sodium carbonate was preferred, removing the remaining Na from the gas, while simul-taneously converting sodium telluride to sodium carbonate. The conversion of telluride tocarbonate continued until the second vertical line was reached. The sodium silicate phasedecreased after the first vertical line.Figure 4.24: Sodium distribution as a function of initial oxygen amount at 1000 ◦C. Fixedinputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.794.8. Effect of Initial OxygenFigure 4.25 shows the distribution of silver as a function of initial oxygen amount at1000 ◦C. It can be seen that the silver present as elemental silver was largely unaffected bythe increase in initial oxygen until the O2 concentration at the first vertical line was reached.As the oxygen continued to increase, the silver was converted to Ag2Te, until a point atwhich all of the Ag was in the form of Ag2Te. This conversion was due to the previouslyexamined conversion of sodium into sodium carbonate. Sodium telluride was found to beconverted to sodium carbonate, which released Te into the system, which formed Ag2Te.As more sodium telluride converted, the concentration of Ag2Te increased. Once all of theAg was in the form of Ag2Te, further increases in oxygen led to a reversal of the conversion,forming elemental Ag, which reformed the metal phase. The sudden removal of Ag2Te isfurther explained in the following figure, which examines the Te distribution.Figure 4.25: Silver distribution as a function of initial oxygen amount at 1000 ◦C. Fixedinputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.804.8. Effect of Initial OxygenFigure 4.26 shows the distribution of tellurium as a function of initial oxygen amountat 1000 ◦C. It can be seen that until the first vertical line, the tellurium was present assodium tellurides. Upon reaching the first threshold, the sodium telluride began to convertto other tellurides such as PbTe and Ag2Te, with some being converted to elemental Te.This continued to occur until the second vertical line threshold was reached, at which pointthe lead telluride and elemental tellurium oxidized, forming both lead and tellurium oxides.Once the PbTe and elemental Te were oxidized, the Ag2Te began oxidizing as well, whichimplied that the oxidation of PbTe and Te was preferential to that of Ag2Te. Unlike PbTeand Te, not all of the Ag2Te was oxidized when the maximum of the oxygen input wasreached.Figure 4.26: Tellurium distribution as a function of initial oxygen amount at 1000 ◦C. Fixedinputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.814.9. Effect of Pre-Roasting4.9 Effect of Pre-RoastingThrough the use of HSC Sim 9, the effect of a pre-roasting step on the pyrometallurgicalsystem was examined. In this pre-roasting simulation, activity coefficients were not able tobe input into the simulation. As a result, the absences of these coefficients means that thesystem assumed an ideal state for all species, which may lead to an over or under reportingof the species. The fixed inputs to the system were 100 tonnes/hour of slag, 10 tonnes/hourof Na2CO3, and 20 tonnes/hour of PbO. The roasting temperatures were varied from 300to 500 ◦C, in increments of 50 ◦C. The furnace temperature refers to the temperature usedduring the pyrometallurgical processing step after the pre-roasting step, and was variedfrom 950 to 1150 ◦C in increments of 10 ◦C.Figures 4.27 and 4.28 show the baseline runs for the Ag/Te and Te/Ag ratios, and theTe and Ag recoveries, where there was no pre-roasting step. Variances between these resultsand the pyrometallurgical model discussed previously in this section can be attributed to alack of activity coefficients.Figure 4.27: Simulation baseline Ag/Te and Te/Ag ratios. Fixed inputs were 100tonnes/hour of slag, 20 tonnes/hour of PbO, and 10 tonnes/hour of Na2CO3.824.9. Effect of Pre-RoastingFigure 4.28: Simulation baseline silver and tellurium recoveries. Fixed inputs were 100tonnes/hour of slag, 20 tonnes/hour of PbO, and 10 tonnes/hour of Na2CO3.834.9. Effect of Pre-RoastingFigure 4.29 shows the deviation from the baseline Ag/Te ratio for various roastingtemperatures as a function of furnace temperature. It can be seen that pre-roasting tem-peratures of 400 ◦C or higher led to an increase in this ratio. This effect was amplifiedas the pre-roasting temperature increased, due to the higher reaction temperatures presenthaving a positive effect Ag/Te ratio of the system, as previously seen in Figure 4.3 up totemperatures of 1016 ◦C.Figure 4.29: Deviation from baseline for the Ag/Te ratio of the metal phase as a function ofroast and furnace temperature. Fixed inputs were 100 tonnes/hour of slag, 20 tonnes/hourof PbO, and 10 tonnes/hour of Na2CO3.844.9. Effect of Pre-RoastingFigure 4.30 shows the deviation from the baseline silver recovery to the metal phasefor various roasting temperatures as a function of furnace temperature. Over the entiretemperature range examined, for both the roast and furnace temperature, the maximumincrease seen in the silver recovery was 0.035%.Figure 4.30: Deviation from baseline for the silver recovery to the metal phase as a func-tion of roast and furnace temperature. Fixed inputs were 100 tonnes/hour of slag, 20tonnes/hour of PbO, and 10 tonnes/hour of Na2CO3.854.9. Effect of Pre-RoastingFigure 4.31 shows the deviation from the baseline Te/Ag ratio in the slag phase forvarious roasting temperatures as a function of furnace temperature. It can be seen thatthe trend here follows the one previous examined in Figure 4.29, the Ag/Te ratio of themetal phase. Pre-roast temperatures above 400 ◦C led to an increase in the ratio, howeverthe increase in the ratio was not as significant as that seen in the Ag/Te deviation frombaseline.Figure 4.31: Deviation from baseline for the Te/Ag ratio of the slag phase as a function ofroast and furnace temperature. Fixed inputs were 100 tonnes/hour of slag, 20 tonnes/hourof PbO, and 10 tonnes/hour of Na2CO3.864.10. Leaching - Eh-pH DiagramFigure 4.32 shows the deviation from the baseline tellurium recovery to the slag phasefor various roasting temperatures as a function of furnace temperature. Similar to the Agrecovery to the metal phase, the largest change occurred at high pre-roasting temperaturesand low furnace temperatures. Unlike the Ag recovery, the change in Te recovery wasextremely small, the largest change being approximately 0.0031%. This implied that apre-roasting step did not aid Te recovery.Figure 4.32: Deviation from baseline for the tellurium recovery to the slag phase as afunction of roast and furnace temperature. Fixed inputs were 100 tonnes/hour of slag, 20tonnes/hour of PbO, and 10 tonnes/hour of Na2CO3.4.10 Leaching - Eh-pH DiagramThe Eh-pH diagram for tellurium at 25 ◦C can be seen in Figure 4.33. This diagram wasgenerated using the Eh-pH module of HSC Chemistry 9.1. Using this diagram, it can beseen that at pH values above 11 (the expected value when leaching), the dominant telluriumspecies was TeO2−3 . Upon lowering the pH of the solution to below 9, the formation of solidTeO2 was expected. This could provide a secondary pathway for the recovery of telluriumvia tellurium dioxide, if the solution chemistry was not conducive to electrowinning.874.11. Tellurium Dioxide PrecipitationFigure 4.33: Eh-pH diagram of tellurium in solution at 25 ◦C.Add silver Eh pH Diagram4.11 Tellurium Dioxide PrecipitationFigure 4.34 shows the fractional distribution of Te species as a function of pH in thesolution. At pH values over 6, several different aqueous tellurium species exist in solution.At pH values lower than 6, TeO2 was expected to precipitate from solution. Further loweringthe pH to values below 1 appeared to reverse the precipitation, where the soluble TeOOH+formed.884.11. Tellurium Dioxide PrecipitationFigure 4.34: Fractional distribution of Te species as a function of pH.89Chapter 5Results and Discussion -PyrometallurgyThis chapter focuses on the high temperature processing of the slag, in order to separatethe silver form the tellurium. The results of the experiments are presented in a statisticalmodel. Analysis of the processed slag via XRD, proposed reactions, and a comparison tothe thermodynamic model are shown.5.1 Preliminary ResultsResults from the preliminary pyrometallurgical tests are reported in the following sec-tions. These results influenced the main CCD pyrometallurgical testing methods and anal-ysis design.5.1.1 Effect of Silica AdditionPreliminary results on analyzing the effect of silica are presented in this section. Thex-values are presented in terms of Q (the silicon/sodium ratio). Silicon was added to thesystem in the form of silica (SiO2). Figure 5.1 shows the Ag and Te grade of the metal phaseas a function of Q. It can be seen that as the Q value increased, the Ag grade decreased,while the Te grade subsequently gradually increased. A maximum Te grade occurred at aQ value of 0.125.905.1. Preliminary ResultsFigure 5.1: Silver and tellurium grade in the metal phase as a function of the silicon tosodium ratio, Q. The temperature was fixed at 1000 ◦C with 10 g of PbO addition. Threereplicates were conducted for each test.915.1. Preliminary ResultsFigure 5.2 shows the Ag and Te recovery to the metal phase as a function of Q. Inter-estingly, unlike the grade, the recovery of Ag increased with addition of silica from a Q of0.100 to 0.125. Increasing the silica concentration led to the reduction in the Ag recovery;however, all the data points corresponding to Q > 0.125 were within each other’s standarddeviations. Tellurium recovery followed a similar trend, where an increase in Q resulted inan initial increase in recovery, followed by a decrease.Figure 5.2: Silver and tellurium recovery to the metal phase as a function of the silicon tosodium ratio, Q. The temperature was fixed at 1000 ◦C with 10 g of PbO addition. Threereplicates were conducted for each test.925.1. Preliminary ResultsFigure 5.3 shows the silver to tellurium ratio as a function of Q. It can be seen that asQ increased, the Ag/Te ratio decreased, as expected due to the increasing Te recovery anddecreasing Ag recovery. The Ag/Te ratio did not appear to change at Q values of 0.125 andhigher, as the values were all within each other’s standard deviations.Figure 5.3: Silver to tellurium ratio in the metal phase as a function of the silicon to sodiumratio, Q. The temperature was fixed at 1000 ◦C with 10 g of PbO addition. Three replicateswere conducted for each test.5.1.2 Effect of Lead Oxide AdditionFigure 5.4 shows the recovery of tellurium and silver to the metal phase as a functionof PbO addition. It can be seen that increasing the amount of PbO led to the increase ofboth Ag and Te recovery to the metal phase. Tellurium recovery was found to decreaseonce 10 g of PbO or more was added. This decrease in Te recovery was unexpected, asthe thermodynamic model predicted that increasing the amount of PbO led to a higher Terecovery, due to the formation of PbTe in the metal phase. The addition of no PbO led toa metal phase not forming, which was why the recovery of both Ag and Te at 0 g PbO was0%. In this case, instead of forming a metal phase below that of the slag, the metal wasdispersed in small beads throughout the slag.935.1. Preliminary ResultsFigure 5.4: Silver and tellurium recovery to the metal phase as a function of PbO addition.The temperature was fixed at 1000 ◦C. Three replicates were conducted for each test.5.1.3 Effect of Sodium Carbonate AdditionSodium carbonate addition was found to have a beneficial effect on the system. With theaddition of Na2CO3 to the system, the recovery of tellurium and silver to the metal phaseincreased from 4.72 to 5.06% (Te) and 28 to 47.5% (Ag). This increase in Ag recovery ledto a slight increase in the Ag/Te ratio from 4.61 to 4.66. This ratio would have been higherif the Te recovery did not increase as well. When combined with an initial pre-roastingstep, as described in Section 5.1.5, the effects were amplified and led to a further increasein the Ag recovery to the metal phase.5.1.4 Effect of Temperature and TimeThe effect of temperature and time were not thoroughly examined in the preliminarytesting. Experiments that varied the time from 30 to 45 minutes while holding all othervariables fixed found that increasing the time led to an increase in silver recovery to themetal phase from 9 to 19%. Tellurium recovery to the metal phase was subsequently foundto increase from 0.26 to 0.91%.945.2. Central Composite Design Results5.1.5 Effect of Pre-RoastTwo identical batches of slag were prepared and ran in triplicate to determine the effectof a pre-roasting step on the Ag and Te behaviour in the system. The samples had a Qvalue of 0.15, Na2CO3 addition of 5 g, and a PbO addition of 15 g. The temperature of thepyrometallurgical processing was 1000 ◦C and samples were held at this temperature for 30minutes. The roast condition was 400 ◦C for 1 hour.Table 5.1 compares the two runs. The values in brackets are the standard deviation.Table 5.1: Comparison of pyro and pre-roast and pyro on several response variables in themetal phase. The values in brackets indicate the standard deviation (σ), three replicateswere conducted per test.Response Variable Pyro Only Pre-Roast + PyroAg Grade (%) 19.04 (1.33) 25.11 (1.19)Ag Recovery (%) 47.47 (2.83) 75.71 (5.71)Te Grade (%) 4.09 (0.19) 0.71 (0.10)Te Recovery (%) 5.06 (0.29) 1.06 (0.11)Ag/Te Ratio 4.56 (0.52) 35.80 (4.98)Metal Mass Recovered (g) 16.37 (0.24) 19.78 (1.17)It can be seen from Table 5.1 that the addition of a pre-roasting step lead to the increasein both the Ag recovery and grade in the metal phase, while the Te recovery and gradedecreased. Both of these effects are desirable as it led to a cleaner metal phase, containingless tellurium, as seen by the increase in the Ag/Te ratio of over 7x. Therefore, the pre-roasting step was determined to be beneficial to the pyrometallurgical processing of the slagand was incorporated into the main series testing.5.2 Central Composite Design ResultsThe results from the central composite design results are provided in this section. Thecenter point runs, which served as an indicator of the error in the runs, are displayed inrelevant figures, along with the statistical modelling results. The results from each of the36 CCD tests can be found in Appendix D. These results were used as the inputs for thestatistical model, which is described further in this section. Any tellurium which did notreport to the metal or slag phase was assumed to be lost to the gas/dust phase.955.2. Central Composite Design Results5.2.1 Center Point ResultsThe response variables for the center point runs are examined in this section. Since theexperimental conditions used for the center point runs were identical, ideally the resultsshould be the same (i.e., with a high level of repeatability).Figure 5.5 shows the tellurium recovery to the slag phase for each center point run. Theaverages are shown by the x-marker, while the standard deviation is shown by the bars. Theoverall mean and standard deviation of the mean are seen in the upper left hand corner. Itcan be seen that the tellurium recovery to the slag phase increased with increasing numberof the center point runs. This was unexpected, as the identical nature of the center pointruns should have led to identical results. One potential explanation for the variation wasthe fact that the runs were conducted over a 2 month period, and changes to the ambientair conditions may have affected the experiments. The continual use of the furnace may alsohave changed the heating profile (due to wear on the heating elements), which could haveaffected the results. It can also be seen that for run 6 and onwards, the trend stabilized,where the mean did not fluctuate as it did in the earlier runs. This could potentially beexplained by the installation of a HEPA vacuum for capturing dusts and gases prior tocenter point run 6. This vacuum, although mounted above the exhaust of the furnace, mayhave induced a negative pressure inside the furnace, leading to the removal of gaseous andairborne dust faster than in the previous tests.965.2. Central Composite Design ResultsFigure 5.5: Tellurium recovery to the slag phase for each center point test. The fixed inputswere the temperature at 1000 ◦C, 400 ◦C pre-roast for 90 minutes, 50 g slag, 10 g PbOaddition, and 5 g Na2CO3 addition. Three replicates were conducted per test.Averaging all the center point runs, the mean tellurium recovery to the slag phase wasobserved to be at least 75% (with the runs after center point 6 stabilizing around 90 to95% recovery). This recovery value was high, however the missing tellurium likely reportedto the dust or gas phase. This implied that a dust or gas capture system would likely benecessary for maximum tellurium recovery on an industrial scale.Figure 5.6 shows the silver recovery to the metal phase for each center point run. Theaverages are shown by the x-marker, while the standard deviation is shown by the bars.The overall mean and standard deviation of the mean are seen in the bottom left handcorner. It can be seen that for the majority of tests, the mean was between 70 and 90%recovery. Center point run 6 was an outlier for this, where the mean recovery was 61.4%.Similar to the Te recovery as seen in the previous figure, the Ag recovery had an upwardtrend as runs progressed. This trend, however, was not as steep as the one seen in the Terecovery.975.2. Central Composite Design ResultsFigure 5.6: Silver recovery to the metal phase for each center point test. The fixed inputswere the temperature at 1000 ◦C, 400 ◦C pre-roast for 90 minutes, 50 g slag, 10 g PbOaddition, and 5 g Na2CO3 addition. Three replicates were conducted per test.It can also be seen that for center point runs 2, 5, and 9, the standard deviation ofthe Ag recovery achieved recoveries above 100%. In these runs, one of the triplicate runsreported a recovery of over 100%. This recovery of over 100% can be due to leftover silverfrom a previous run, where the metal phase was not completely poured out, leaving behindsome Ag. Additionally, during the pour, if some slag were to somehow mix in with themetal phase and not be removed, then the reported weight of the metal phase would behigher, which led to a higher reported silver recovery due to the mass balance method whichwas used. Lastly, if the input slag for that run contained less silver than usual, despite bestefforts to ensure a homogenous input, or if not enough input slag was added due to operatorerror while measuring out the slag, then less Ag would be present in the input material thanexpected, leading to higher recoveries.Figure 5.7 shows the silver grade of the metal phase for each center point run. Theaverages are shown by the x-marker, while the standard deviation is shown by the bars. Theoverall mean and standard deviation of the mean are seen in the bottom left hand corner.The silver grade of the metal phase was consistent between runs, varying by approximately10% from the mean. The remaining grade was typically between 30 and 50% Pb, 10-15%985.2. Central Composite Design ResultsBi, 1% Cu, and a low percentage (<1.5%) impurity. The grade followed the trends seenwith the silver recovery, where a lower recovery led to a lower grade.Figure 5.7: Silver grade of the metal phase for each center point test. The fixed inputs werethe temperature at 1000 ◦C, 400 ◦C pre-roast for 90 minutes, 50 g slag, 10 g PbO addition,and 5 g Na2CO3 addition. Three replicates were conducted per test.Figure 5.8 shows the total impurity grade of the metal phase for each center point run.The averages are shown by the x-marker, while the standard deviation is shown by thebars. The overall mean and standard deviation of the mean are seen in the bottom righthand corner. The impurity elements selected were aluminum, arsenic, magnesium, sodium,selenium, antimony, silicon, and tellurium. It can be seen that the mean impurity grade forcenter point runs other than the center run was between 1 and 1.5%.995.2. Central Composite Design ResultsFigure 5.8: Total impurity grade of the metal phase for each center point test. The fixedinputs were the temperature at 1000 ◦C, 400 ◦C pre-roast for 90 minutes, 50 g slag, 10 gPbO addition, and 5 g Na2CO3 addition. Three replicates were conducted per test.The mean of the first center point run was 2.36%, approximately 1% higher than themean seen in the following runs. This higher impurity mean was likely due to the factthat the center point one was the first test conducted, and as such there may have beenerrors in the process which were removed in subsequent tests. The fact that the followingtests means appear to be similar indicates that any errors were mitigated or removed insubsequent experiments.5.2.2 Statistical Modelling ResultsThe results for four statistical models generated in MATLAB via the stepwiselm func-tion are presented in Table 5.2. The full quadratic models generated using fitlm aredisplayed in Appendix F. The input terms were all normalized prior to input, as describedin Section 3.6.2, which allowed for direct comparison between terms. It can be seen thatfor each model produced, the X1 - Temperature, X3 - PbO, X4 - Na2CO3, and X5 - RoastTemperature terms were present, while X2 - Time was not seen in the tellurium recovery tothe slag phase model, while being present in the other three models. This implies that atthe selected levels (Padd and Premove values) of statistical modelling, the effect of time did1005.2. Central Composite Design Resultsnot show a statistically significant effect on the tellurium recovery. However, the telluriumslag recovery model itself did not display satisfactory statistical significance, as per thestatistical values seen at the bottom of the table. For the Te slag recovery model, the R2and R2adj terms were very low, with the R2adj value being negative. These values imply thatthe model did not predict the variance in the model well. In addition to this, the F -testvs. a constant model (i.e. one intercept term) was low, at 0.888, meaning that the modelwas only slightly better than a one term model at predicting the Te slag recovery. Lastly,the p-value for the model was 0.529, which means that the model failed at many commonlyseen statistical levels (α = 0.95 for example). This large p-value suggests that the selectedinput variables may not have a statistically significant effect on the tellurium recovery tothe slag phase. However, due to the complexity of the slag system, it was likely that otherinteractions, not accounted for in this model, may have had an effect on the system, andas such a different or more complex statistical model may be used to better predict thesystem. The other three models are seen to have higher values for R2, R2adj , and the F -testvs constant model. In each of the three models, the R2adj values are all over 0.6, whichmeans that the model cannot predict 40% or less of the variability seen. Furthermore, thep-values of these models all imply that the input variables have a statistically significanteffect on the system.1015.2. Central Composite Design ResultsTable 5.2: Pyrometallurgical CCD testing statistical model output.Coefficients TeSlagRec TeMetalRec AgMetalRec ImpurityMetalGrConstant 88.55 0.2036 78.98 1.592LinearX1 - Temperature 0.759 -0.5303 6.373 -0.6144X2 - Time -0.1546 -0.2600 -0.1572X3 - PbO 0.1136 0.2573 9.051 -0.1423X4 - Na2CO3 -0.08378 0.2019 4.470 0.1192X5 - Roast Temperature 0.6138 0.1407 1.643 0.1215InteractionX1X2 0.1676X1X3 1.000 -0.2769X1X4 -0.2652 -0.2734X1X5 -0.1992 -0.2364X2X4 -0.1726X2X5 -0.2304 4.349 -0.1963X3X4 -1.056 0.2306 0.2342X3X5 -1.268 0.1682SquareX12 0.2488 -3.514 0.3050X22 0.1267X32 -4.277StatisticsPRemove 0.30 0.05 0.1 0.05PAdd 0.15 0.025 0.05 0.025R2 0.176 0.904 0.698 0.823R2adj -0.0233 0.836 0.612 0.755F vs Const. Model 0.888 13.2 8.1 12.1Model p-value 0.529 1.78e-07 0.0013 1.91e-07The statistical results of the model led to the conclusion that the tellurium recovery tothe slag phase cannot be accurately predicted using this model. However, the statisticalsignificance of the other three models, all pertaining to the metal phase, imply that themodel can be used for prediction of the metal phase. This allows for examining the system1025.2. Central Composite Design Resultsfrom the perspective of tellurium rejection from the metal phase as opposed to the previousviewpoint of tellurium recovery to the slag phase. Minimization of the tellurium recoveryto the metal phase can be used as an objective of any optimization to be done, since anytellurium not present in the metal must report to one of the other two phases, the slag or thedust/gas phase. If in the dust/gas phase, the use of a scrubber or other gas capture systemcan prove effective in recovery of the tellurium, and any tellurium which reports to the slagcan be recovered through the processes examined in future chapters of this dissertation.The following subsections examine the four models seen in Table 5.2. The figures pre-sented show the predicted recovery or grade, and act as a visual indicator of the magnitudeof the coefficients seen in the previously mentioned table. Since there were five input vari-ables and only two are compared in each figure, the remaining input variables were set totheir baseline values as follows: X1 - Temperature 1000 ◦C, X2 - Time 45 minutes, X3 -PbO 10 g/50 g slag, X4 - Na2CO3 5 g/50 g slag, X5 - Roasting Temperature 400◦C. Thesevalues were selected as they were the midpoint of their respective examined ranges.Tellurium Recovery to the Slag PhaseFigure 5.9 shows the observed versus predicted values for the tellurium recovery to theslag phases. The predicted values were determined using the equation produced from thestatistical modelling, the coefficients of which where are displayed in Table 5.2. The dashedline indicates a perfect correlation. The figure, in conjunction with the R2 and R2adj valuespresented in the previous table, shows a poor fit of the model to the experimental data.The data points show little correlation to an overall trend, and while there does appear tobe an upwards trend visible (loosely following the dashed line) another, horizontal trendcan also be seen to exist at a predicted Te recovery of 86.3%. This horizontal trend is avisual indication of the low F -statistic (0.888) compared to a constant model.1035.2. Central Composite Design ResultsFigure 5.9: Observed versus predicted tellurium recovery values to the slag phase. Thedashed line indicates a perfect correlation.Tellurium Recovery to the Metal PhaseFigure 5.10 shows the observed versus predicted values for the tellurium recovery to themetal phases. The predicted values were determined using the equation produced from thestatistical modelling, the coefficients of which are displayed in Table 5.2. The dashed lineindicates a perfect correlation. It can be seen that model fit was good. The majority ofthe points were seen to follow the trendline, as indicated by the high R2 and R2adj values.However, one negative aspect of the model was that some of the predicted values can beseen to be negative, which can obviously not happen. Future iterations of the model willneed restrictions to ensure that values cannot be less than zero.1045.2. Central Composite Design ResultsFigure 5.10: Observed versus predicted tellurium recovery values to the metal phase. Thedashed line indicates a perfect correlation.Figure 5.11 shows the predicted tellurium recovery to the metal phase as a function oftemperature and time. Due to the two square terms present for both X1 - Temperatureand X2 - Time, a minimum occurred in the plot, unlike all the saddle point graphs seenpreviously. These two terms are the dominant terms due to them appearing in linear,interaction, and square coefficients. The X1 - Temperature term was seen to have thelargest absolute value of the linear terms, with several of its interaction terms having largeabsolute values as well. The X1 - Temperature square term was nearly double that of theX2 - Time term, seen in the graph by the oblong shape of the contour line around themaximum.1055.2. Central Composite Design ResultsFigure 5.11: Predicted tellurium recovery to the metal phase as a function of temperatureand time. Fixed inputs were 10 g of PbO / 50 g slag, 5 g Na2CO3 / 50 g slag, and apre-roasting temperature of 400 ◦C.Figures 5.12 and 5.13 show the tellurium recovery to the metal phase as a function oftime and either sodium carbonate or roasting temperature. In both of these models, thevalue of the coefficients for X4 - Na2CO3 addition and X5 - Roasting Temp. were similar.The trends for both of these plots were similar, where the effect of roasting temperaturehaving a larger influence on the Te recovery. For shorter durations, the addition of moreNa2CO3 or increasing the roasting temperature led to a higher tellurium recovery to themetal phase, with the roasting temperature having a larger impact and leading to higherrecoveries (3.1% compared to 2.9%). While this change appeared small, any increase in theTe recovery to the metal phase results in decreased purity of the metal. Once the durationwas increased beyond approximately 60 minutes, the increase in either X4 - Na2CO3 or X5- Roasting Temp. led to a decrease in the Te recovery.1065.2. Central Composite Design ResultsFigure 5.12: Predicted tellurium recovery to the metal phase as a function of time andNa2CO3 addition. Fixed inputs were 10 g of PbO / 50 g slag, 1000◦C pyrometallurgicalprocessing temperature, and a pre-roasting temperature of 400 ◦C.1075.2. Central Composite Design ResultsFigure 5.13: Predicted tellurium recovery to the metal phase as a function of time androasting temperature. Fixed inputs were 10 g of PbO / 50 g slag, 5 g of Na2CO3 / 50 gslag and 1000 ◦C pyrometallurgical processing temperature.Figure 5.14 shows the recovery of tellurium to the metal phase as a function of leadoxide and sodium carbonate additions. It can be seen that the increase of either one ofthese variables led to a decrease in the Te recovery, whereas increasing both simultaneouslyincreased the tellurium recovery. The linear and interaction terms X3 - PbO, X4 - Na2CO3,X3X4 were all positive, which led to this increase. However, the other interaction termsfeaturing X3 - PbO and X4 - Na2CO3 were largely negative, which meant that increasingthe other variables beyond the baseline led to a decrease in the recovery.1085.2. Central Composite Design ResultsFigure 5.14: Predicted tellurium recovery to the metal phase as a function of PbO andNa2CO3 additions. Fixed inputs were 1000◦C pyrometallurgical processing temperaturefor 45 minutes, a pre-roasting temperature of 400 ◦C.Overall, the tellurium recovery to the metal phase model appears statistically significantwith the high R2, R2adj values, and F -statistic, and the low p-value. Thus, this model canbe used to help minimize the tellurium recovery to the metal phase, forcing it to reportto the slag or dust, from which it can be collected and recovered. The downside to thismodel was that fact that negative values could exist in the model, which is impossible inreality. The negative terms exist due to the function generated by the model being able togenerate these values as an underestimation of the Te recovery. This was due to the inputdata set values being close to zero, so as a result, the quadratic fit extended below zeroin an attempt to maximize the fit. As such, care must be taken when using this model toensure that negative values are not obtained, and that the model use is within the range ofvariables that cannot result in negative recoveries.Silver Recovery to the Metal PhaseFigure 5.15 shows the observed versus predicted values for the silver recovery to themetal phases. The predicted values were determined using the equation produced from the1095.2. Central Composite Design Resultsstatistical modelling, the coefficients of which are displayed in Table 5.2. The dashed lineindicates a perfect correlation. It can be seen that the data points loosely follow the trend,as reflected in the R2 and R2adj values of 0.698 and 0.612, respectively. These values implythat the model cannot account for approximately 30 to 40% of the variance in the model.Figure 5.15: Observed versus predicted silver recovery values to the metal phase. Thedashed line indicates a perfect correlation.Figure 5.16 shows the silver recovery to the metal phase as a function of temperatureand lead oxide addition. The X1 - Temperature and X3 - PbO terms can be stated tobe the most significant for the system, due to the fact that these two linear terms wherethe largest, and that the only two square terms in the model were for these terms. Thesquare terms are responsible for the formation of a maximum point in the figure, seen ata temperature of 1040 ◦C and 14 g PbO addition. The magnitude of the linear terms canbe said to ‘flatten’ the top of this curve, which was apparent by the slope in the figure,indicated by the distance between contour lines. It can be seen that with distance awayfrom the center (14 g PbO addition), the distance between the contours decreased, owningto the square terms increasing in magnitude. Close to the maximum, the linear termsexerted more influence and acted to flatten the peak.1105.2. Central Composite Design ResultsFigure 5.16: Predicted silver recovery to the metal phase as a function of temperature andPbO addition. Fixed inputs were 45 minutes processing time, 5 g of Na2CO3 / 50 g slag,and a pre-roasting temperature of 400 ◦C.Figure 5.17 shows the recovery of silver to the metal phase as a function of lead oxideand sodium carbonate additions. Due to the square term seen for X3 - PbO, a paraboliccylindrical profile can be observed in the figure. It can be seen that the addition of anyNa2CO3 led to an increase in the silver recovery. This was due to there only being onepositive linear term for X4 - Na2CO3, so any increase in X4 - Na2CO3 translated to anincrease in Ag recovery. For X3 - PbO, an optimum value of approximately 14 g additionled to the highest Ag recovery across all Na2CO3 additions. Values lower or higher than 14g PbO per 50 g slag led to a decrease in the Ag recovery.1115.2. Central Composite Design ResultsFigure 5.17: Predicted silver recovery to the metal phase as a function of PbO and Na2CO3additions. Fixed inputs were 1000 ◦C pyrometallurgical processing temperature for 45minutes and a pre-roasting temperature of 400 ◦C.Figure 5.18 shows the silver recovery to the metal phase as a function of roasting temper-ature and time. The only interaction term present in the model was X2X5, an interactionbetween the time and the roasting temperature. In their linear terms, X2 - Time was seento be the smallest absolute term and the only one which was negative, while X5 - RoastingTemp. was the second smallest absolute term. The positive interaction coefficient betweenthese terms led to an increase in the Ag recovery when both terms were increased simulta-neously above the baseline value. However, at the maximum values for both of these terms,the recovery was predicted at 108%, which is not possible.1125.2. Central Composite Design ResultsFigure 5.18: Predicted silver recovery to the metal phase as a function of time and roastingtemperature. Fixed inputs were 10 g of PbO / 50 g slag, 5 g of Na2CO3 / 50 g slag, and apre-roasting temperature of 400 ◦C.Overall, the model describing the silver recovery to the metal phase had an adequatefit, with an R2adj value of 0.612. The F -statistic of 8.1 and a p-value of 0.0013 corroboratethis claim. While the fit was adequate, further refinement may be necessary. However, theaddition of subsequent terms will further reduce the R2adj . Performing optimization andfitting the data to a cubic or custom model may produce a better fit. Temperature andPbO addition were found to be the dominant factors in the recovery of Ag to the metalphase.Total Impurity Grade of the Metal PhaseFigure 5.19 shows the observed versus predicted values for the total impurity grade ofthe metal phases. The predicted values were determined using the equation produced fromthe statistical modelling, the coefficients of which are displayed in Table 5.2. The dashedline indicates a perfect correlation. It can be seen that the data points have an acceptablefit to the trendline. The low p-value implies that at least one of the terms used in the modelhad a statistical impact on the model, whereas the F -statistic, the second highest in the1135.2. Central Composite Design Resultsfour models, implies that this model was better than a one term model.Figure 5.19: Observed versus predicted impurity grade of the metal phase. The dashed lineindicates a perfect correlation.Figure 5.20 shows the total impurity grade as a function of temperature and sodiumcarbonate addition. X1 - Temperature had the only square term present in the model,and can be seen in the figure by the parabolic shape obtained as the temperature increasesat a constant Na2CO3 addition. Temperature had a negative linear and interaction term,which led to a fairly ‘flat’ area in the range of approximately 980 to 1060 ◦C. When thetemperature increased or decreased further beyond these temperatures, the square termwould become dominant and led to larger changes in the impurity grade. The X4 - Na2CO3had positive linear and interaction terms with the exception of the X1X4 interaction term,which was negative. This led to the total impurity grade increasing as more Na2CO3 wasadded at the lower temperature ranges. This effect reversed at higher temperatures, above1040 ◦C, where the addition of Na2CO3 led to a reduction in the impurity grade.1145.2. Central Composite Design ResultsFigure 5.20: Predicted impurity grade of the metal phase as a function of temperature andNa2CO3 addition. Fixed inputs were 45 minutes processing time, 10 g of PbO / 50 g slag,and a pre-roasting temperature of 400 ◦C.Figure 5.21 shows the total impurity grade of the metal phase as a function of roasttemperature and time. X2 - Time had a negative linear and interaction term X2X5 (Time -Roasting Temp.), which led to a decrease in the impurity grade at values above the baselinevalues of 45 minutes and 400 ◦C. X5 - Roast Temperature had a positive linear term, butnegative interaction terms, which led to the effect seen in the figure, where times of 50minutes or shorter led to higher impurity grades of the metal phase as roast temperaturesincreased, and above 50 minutes the opposite effect occurred.1155.2. Central Composite Design ResultsFigure 5.21: Predicted impurity grade of the metal phase as a function of time and roastingtemperature. Fixed inputs were 1000 ◦C pyrometallurgical processing temperature, 10 g ofPbO / 50 g slag, and 5 g of Na2CO3 / 50 g slag.Figure 5.22 shows the total impurity grade of the metal phase as a function of lead oxideand sodium carbonate additions. It can be seen that increasing both of the inputs simulta-neously led to a saddlepoint at 6 g Na2CO3 and 8 g PbO additions. Further simultaneousincreases led to a larger impurity grade. At the extreme additions of either species, the ad-dition of the other species led to a higher impurity grade. X4 - Na2CO3 likely had a largerinfluence on the system, as it contained one linear and two interaction terms, compared toone linear and one interaction term for X3 - PbO. While the linear values for X4 - Na2CO3were smaller in absolute value than X3 - PbO, the inclusion of the additional interactionterms allowed it greater influence on the system.1165.2. Central Composite Design ResultsFigure 5.22: Predicted impurity grade of the metal phase as a function of PbO and Na2CO3additions. Fixed inputs were 1000 ◦C pyrometallurgical processing temperature for 45minutes and a pre-roasting temperature of 400 ◦C.Overall, the model predicting the impurity grade of the metal phase was found to beadequate. The R2adj value implies that only 24.5% of the variance cannot be explained bythe model. This was the second highest R2adj value obtained. Similarly, the F -statistic wasthe second highest value observed for all the models that were developed. These statisticalvalues imply that the model adequately explains the system, however further improvementshould be possible with different models or additional testing. Temperature was found tohave the most impact on the output of the model, with the sodium carbonate and roasttemperature having the second highest effect (both having the second most terms andsimilar coefficient magnitudes).5.2.3 Distribution of Other ElementsIn addition to silver and tellurium, several other elements were present in the systemand their distributions were examined.1175.2. Central Composite Design ResultsAluminumAcross all tests, the average recovery of aluminum to the slag phase was 170%. Thisvalue was due to the crucible (comprised of glazed alumina) being dissolved over the courseof the runs. When fit to a linear model with Premove and Padd values of 0.25 and 0.15, amodel with an R2adj value of 0.314 was produced. This R2adj value was higher than thatof the Te model produced, implying that the model can predict the recovery of aluminumto the slag phase better than it can for Te. The F -statistic versus a constant model wasfound to be 2.65 and the p-value was 0.0224, which implied that one term in the modelhad some statistical significance at the 97.5% confidence level. This value was likely thetemperature, which had a linear, square, and 2 interaction terms, all of which were positive,which implies that increasing the temperature will greatly increase the aluminum recoveryto the slag phase via the dissolution of the crucible into the slag.ArsenicSimilar to tellurium, modelling the arsenic recovery to the metal phase produced astatistically better model compared to modelling the As recovery to the slag phase (R2adj0.702 Recmetal, 0.348 Recslag; F vs const. model 7.63 (metal), 2.72 (slag)). The averagerecovery to the metal phase was 37% across all runs, with the temperature, time, and sodiumcarbonate addition being identified as the most significant inputs, all having positive squareterms which led to the increase in As recovery as their magnitudes increased.BismuthBismuth recovery to the metal phase had an average of 85% across all runs. When astatistical model was produced for the bismuth recovery to the metal phase a R2adj value of0.687 was obtained with an F -statistic of 7.57 and a p-value below 1.0e-4. The dominantterms were temperature and PbO addition, each with square terms in the model. Tem-perature terms were largely negative, implying a decrease in recovery as the temperatureincreased. PbO addition had a positive linear and X1X3 (Temperature - PbO) interactionterm, but a negative square term. As such, at larger additions of PbO, the Bi recovery wasexpected to decrease.CopperCopper was found to have an average recovery to the metal phase of 28%, with theremainder reporting to the slag phase. The statistical model predicted the recovery ofcopper to the slag phase adequately, with a R2adj value of 0.5 and F -statistic of 5.49. The1185.2. Central Composite Design Resultsmodel for the recovery of copper to the metal phase was statistically better than the silvermodel seen in the previous section (Section 5.2.2), having a R2adj value of 0.68 and F -statisticof 9.49. In the slag model, the dominant terms were time, PbO, and Na2CO3 addition, allof which had linear and square terms. Temperature had a negative linear and positiveinteraction term, the only interaction term present in the model.Interestingly, in the metal recovery model, the temperature played a much larger role,having a square term present alongside time and Na2CO3 addition. Again, only one inter-action term, X2X5 (Time - Roasting Temp.), was present. This implies that at this level ofstatistical analysis, there was not much interaction between the terms with respect to thecopper recovery. Time and Na2CO3 addition were found to be the dominant terms over-all. When either of them increased, the recovery of copper to the metal phase decreased,reporting instead to the slag phase.MagnesiumThe magnesium recovery to the slag phase was found to have an average of 103% acrossall runs. Similar to Al, this excessive Mg was due to leaching of Mg from the crucible intothe slag. Recovery of Mg to the metal phase was found to have a maximum value in all runsof 0.7%, with an average value of 0.06%. Due to this low recovery to the metal phase, itwas assumed that Mg recovery to the metal phase was not a concern in the current testing.LeadLead behaviour in the system followed a predictable trend; when the amount of PbOaddition increased so too did the grade of lead in both the silver and slag. Interestingly,when comparing the PbO addition to the lead recovery to the slag phase, no trend wasobserved. Lead recovery was, on average, 5% to the slag phase across all runs, with the runhaving 0 g PbO addition having a higher Pb recovery to the slag phase, 13.2%, comparedto the test where 20 g PbO was added, at 9%. The majority of the remaining lead wasrecovered to the slag phase, with some losses (5%) being present. This loss was likely dueto lead remaining in the crucibles after the run. Losses to the gas or dust phase were alsopossible.AntimonyThe low antimony grade of the input slag (0.55%, as seen in Section 3.2) made theanalysis of Sb recovery difficult due to the quantities involved (0.275 g typical input).Statistical models produced for Sb for both the recovery to the metal and slag phases had1195.2. Central Composite Design ResultsR2adj values of 0.328 (metal) and 0.412 (slag), with F -statistics of 3.2 (metal) and 3.7 (slag).These values are statistically poor, and as such should not be used to infer the behaviour ofSb from the model. Experimental results using ICP and EDS analysis showed the recoveryof Sb to the slag phase to be 15% on average, with the remainder reporting to the metalphase (via EDS analysis). Directly comparing the input with the Sb slag recovery did notreveal any particular correlation between the input variables and the recovery.SeleniumSelenium, similarly to antimony, was present in the system in small quantities (0.3%input grade, 0.15 g typical input), so as such the values examined were more likely tobe subject to error due to rounding issues or measuring error. However, using ICP-OESanalysis the recovery of Se to the slag phase was found to be on average 90%, with nocorrelations found between the input variables and the recovery to the slag phase. ICP-OES analysis of the metal found virtually no Se present in the metal (maximum grade of0.00008% observed), so the remainder of the Se was either lost to the dust or gas.5.2.4 Model Applications and OptimizationThe statistical models produced serve as a starting point for future work, where theoutputs can be predicted prior to the start of laboratory testing. The framework for de-signing the models, the central composite design methodology, can also be applied whenscaling up the system. The lessons learned when conducting the small-scale experimentscan be applied when increasing the volume of the system, which can allow for more accurateresults. In addition, the comparison of coefficients and terms between a laboratory scale,and scaled up test can potentially be used to predict conditions or responses in even largersystems. This model can only be used for small scale processes involving this exact typeof slag, any potential changes in the slag chemistry will lead to deviation from the model.Should different slags be examined in the future, additional terms can be added for themodel in order to account for variance in chemistry. Alternatively, the input terms can benormalized to be in terms of input per element amount (i.e. 5 g PbO / 10 g Te in slag).The equations produced from the models can also be used as objective functions for sys-tem optimization. The equations can also be combined to form a single objective function,the relative weight or importance of different variables can be adjusted by multiplying theoriginal functions by a value. Through this method of optimization, ideal values for theinput variables can be determined for a desired recovery and/or grade target.1205.3. Microstructure Characterization5.3 Microstructure CharacterizationFigure 5.23 shows a BSE SEM image of a polished section of the metal phase producedduring a center point run. It can be seen that there were two distinct phases, a grey,lead-rich matrix, and a black, silver-rich dendrite phase. The grey matrix was comprisedof 80.08% Pb, 16.18% Bi, and 1.33% Cu. Na, As, Ag, Sb, and Te were present in traceconcentrations. The black, silver phase contained 91.57% Ag, 3.80% Pb, 2.67% Sb, 1.44%Cu, 0.29% Te, and 0.23% Bi. The separation of the two phases can be attributed to themiscibility gap seen in the Ag-Pb phase diagram in Figure A.2.Figure 5.23: BSE wide view of the metal phase showing dendrite structures.Figure 5.24 shows a closer view of the dendrite structure, with EDS elemental mapsfor Ag, Bi, and Pb found below the BSE image. The elemental maps clearly show theseparation of Ag from Pb and Bi.1215.3. Microstructure CharacterizationBSEAg Bi PbFigure 5.24: BSE image and elemental mapping of the Ag-Bi-Pb alloy. Processing conditionswere 30 minutes at 1000 ◦C, with additions of 10 g of PbO and 5 g of Na2CO3 per 50 g ofslag, and pre-roasting of 90 minutes at 400 ◦C.Figure 5.25 shows a BSE image of the exterior surface of a crucible used for pyrometal-lurgical testing. The dark phase to the right of the image is the epoxy resin, and as such canbe ignored. From the image, the right side of the crucible appears different when comparedto the center. This is the glazing, and when examined using EDS, it was found to contain1225.3. Microstructure Characterizationa higher Si/Al ratio compared to the center. This exterior was found to contain 32% Si,7% Na and Al, 4.5% Ca, 3.5% K, 1.4% Mg, with the remainder being oxygen. The interior(i.e. crucible material), contained 28% Si, 26% Al, 2% Na, K, and Ca, with the remainderbeing oxygen. Compared the the surface, the interior of the crucible was also found to besubstantially more porous, as seen in the image.Figure 5.25: BSE image of the outside surface of a crucible used for pyrometallurgicaltesting.Figure 5.26 shows a BSE image of the slag-crucible interface from a crucible used forpyrometallurgical testing. Over the course of the experiments, several crucibles were used,typically being replaced after being used for a certain amount of runs (approximately 10),or until a physical defect was observed. These crucibles would be removed from use, toavoid the potential for breakage during a run. Figure 5.26 is from one such crucible, wherea crack was seen on the sidewall after a run. This crack likely developed either due to thermalstresses that evolved between the hot crucible and colder tongs during handling operations,or through excess pressure when gripping the crucible, creating a stress fracture. However,three distinct phases can be seen in the image. The first, bright phase, was determined via1235.3. Microstructure CharacterizationEDS to be a high-tellurium content phase, containing 72% Te, 11.4% Na, and the remainderoxygen. The dark grey matrix around the bright phase was found to have a compositionsimilar to that of both the slag and the crucible material. This matrix was assumed tobe the remnants of the glazing seen in the previous image, or at least a section where theglazing had been dissolved away. EDS assays showed a composition of 21.2% Na, 17.3% Al,15.2% Si, 1.4% Te, 0.6% Ag, with the remainder oxygen. When looking at the image, aninterface can be seen between the two darker grey phases, running diagonally through theimage. EDS assays of the third phase above this line revealed that it had a compositionsimilar to that of the inner crucible material, but with lower amounts of Si (21 compared to28%), and a higher sodium content (5 compared to 2%). The lower Si grade suggests thatleaching of the Si into the slag occurred, whereas the increased Na grade suggests eitherleaching of the Na into the crucible, or a contamination during the polishing process.Figure 5.26: BSE image of the slag-crucible interface from a crucible used for pyrometal-lurgical testing.1245.4. Thermodynamic Model Comparison5.4 Thermodynamic Model ComparisonThe results of the CCD experiments were compared to the results generated by thethermodynamic model as seen in Chapter 4. This comparison could be used to see wherepotential deviations from the thermodynamic model occurred, and would indicate how wellthe thermodynamic model was able to predict the outcome of the experimental system.Figure 5.27 shows the difference in silver recovery to the metal phase compared to thepredicted silver recovery. The y-axis shows the actual results minus the theoretical results,so a negative value indicates how far below the theoretical value the actual test was. Itcan be seen that the majority of tests (all besides three) had negative Δ recovery values,indicating that the actual value was lower than the predicted one. The two runs withpositive Δ recovery values were calculated to have Ag recoveries of over 100%, which is notpossible. This discrepancy was previously explained in Section 5.2.1.Figure 5.27: Experimental to model comparison showing the silver recovery to the metalphase. The y-axis shows the actual minus predicted results.The data point symbols in the figure indicate the time of the actual run. Note, that timecannot be replicated in the thermodynamic model. The differentiation based on time was1255.4. Thermodynamic Model Comparisondone to see if any particular time led to better results than others. Observing the results,there was no apparent trend visible to indicate if one was better than the rest. Each timepoint with multiple runs (everything except 15 and 75 minutes) showed high variability inthe Δ recovery, which implies that the time may play a large role in the process, and thatnone of the times examined allowed the system to reach the equilibrium state as describedby the model.Figure 5.28 shows the difference in tellurium recovery to the slag phase compared tothe predicted tellurium recovery. The y-axis shows the actual results minus the theoreticalresults, so a negative value indicates how far below the theoretical value the actual test was.It can be seen that the predicted Te recovery to the slag phase was over 98%, where theactual recoveries for all tests, besides one, were observed to typically be between 5% and20% lower. The one test with a positive Δ recovery value carried an experimental error,since this would imply a Te recovery to the slag phase of over 100%. What likely occurredwas an error in measuring the amount of slag entering the system, leading to a lower thanexpected amount of Te present. The Δ recovery values for Te compared to the Ag valuesseen in the previous figure were closer to 0, with a maximum Δ of -22.5% for Te, comparedto -66% for Ag. This implies that the thermodynamic model is capable of more accuratelypredicting the Te recovery compared to the Ag recovery. While Δ values of 0 are ideal,further refinement of the model would be necessary to obtain those values.1265.5. Processed Slag Analysis and Proposed ReactionsFigure 5.28: Experimental to model comparison showing the tellurium recovery to the slagphase. The y-axis shows the actual minus predicted results.5.5 Processed Slag Analysis and Proposed ReactionsAn analysis of the processed slag for the star point and center point runs are presentedhere. These star points, seen in Section 3.6.1, Table 3.9, are indicated by the values of -2 or+2, are the extreme points in the model, where between the -2, 0, and +2 states, only onevariable value was changed. These points provide an insight into the effect of each variableon the system.In this section, XRD plots of the -2, 0, and +2 slags for temperature, time, PbO addition,and Na2CO3 addition are examined. Thermodynamic plots showing the species distributionof relevant Te or Ag species are also shown with varying temperature, PbO, and Na2CO3inputs to compare the experimental and theoretical results.An assumption made when analysing the XRD results was that the intensity of thepeaks correlated with the relative amounts of the species present. This assumption wasmade in an attempt to determine potential reactions, however the assumption may carryuncertainty, since only a single XRD scan was done for each run.1275.5. Processed Slag Analysis and Proposed Reactions5.5.1 Effect of TemperatureFigure 5.29 shows XRD spectra comparing tests conducted at 900, 1000, and 1100 ◦C.From the figure, the species determined to change between runs were tellurium, sodiumtelluride, tellurium dioxide, and silver telluride.Figure 5.29: XRD comparison of tests performed at temperatures of 900, 1000, and 1100◦C. Fixed inputs were a processing time of 45 minutes, 10 g of PbO / 50 g slag, 5 g ofNa2CO3 / 50 g slag, and a pre-roasting temperature of 400◦C.Of the three tellurium input species previously stated, only elemental tellurium andtellurium dioxide are present in the XRD spectra. TeO2 also appears to mostly be absentin the 900 ◦C plot, reappearing at 1000 ◦C, before disappearing again. While the removal,formation, and subsequent removal of TeO2 was possible as the temperature increased, alikelier explanation was that the XRD scan did not pick up the TeO2 signal.TeO2 + 2C = Te + 2CO (5.1)TeO2 + Na2O = Na2Te + 1.5O2 (5.2)Na2TeO3 + 3C = Na2Te + 3CO (5.3)1285.5. Processed Slag Analysis and Proposed ReactionsIt can be seen that in the 900 and 1000 ◦C runs, peaks for both Te and Na2Te werepresent in similar ratios, however in the run at 1100 ◦C, the Te peak was observed tosubstaintially shrink, and the Na2Te peak grew in size. This implies that as the temperatureof the system rose above 1000 ◦C, the formation of Na2Te was preferred over that of Te. Tepresent at 1000 ◦C was also likely converted to Na2Te. The following reaction was proposedto occur at these elevated temperatures:Te + Na2O = Na2Te + 1/2O2 (5.4)The removal of oxygen from the system via release into the atmosphere or oxidationof other species will also serve as way for the above reaction to be forced to react in thewritten direction. It was possible that other reactions therefore influence the formation ofNa2Te.When examining the Te recovery to the slag, the tests at 900 and 1100 ◦C had similarTe recoveries, of 88.7 and 88.1%, respectively. It was found that the center point runfeatured the highest Te recovery at 94.7%. The low recovery at 900 ◦C was likely due tohigher amounts of tellurium reporting to the metal phase. As the temperature increased,the kinetics of the process increased, which allowed for the slag to react with the metal,removing tellurium from it. Lastly, at 1100 ◦C, the tellurium was likely lost as dust or gas,as more species were able to volatilize at this temperature.Looking at the behaviour of silver, the recovery to the slag phase was found to decreaseas the temperature went up, which is desirable. Recoveries of 71.0, 26.2, and 9.5% Ag wereobtained for 900, 1000, and 1100 ◦C, respectively. From the XRD analysis, the primary Agphase was found to be Ag2Te, which appeared at 900 and 1000◦C, but was not observedat 1100 ◦C. The proposed reactions involving Ag2Te are listed below:Ag2Te + Na2O = 2Ag + Na2Te + 1/2O2 (5.5)Ag2Te + 2Na = 2Ag + Na2Te (5.6)5.5.2 Effect of TimeFigure 5.30 shows the XRD spectra comparing tests conducted for 15, 45, and 75 min-utes. From the figure, the species of importance were makatite (a sodium silicate species,Na2Si4O8(OH)2·4H2O), tellurium dioxide, silver telluride, and silver tellurite.1295.5. Processed Slag Analysis and Proposed ReactionsFigure 5.30: XRD comparison of tests performed for at 15, 45, and 75 minutes. Fixedinputs were 1000 ◦C pyrometallurgical processing temperature, 10 g of PbO / 50 g slag, 5g of Na2CO3 / 50 g slag, and a pre-roasting temperature of 400◦C.The makatite was only present in the 15 minute run, and likely dissociated into othersilicate species over the course of the run.Regarding tellurium species, the tellurium dioxide was present in the 15 and 45 minuteruns, while not appearing in the 75 minute run. The TeO2 likely reacts with carbon over theduration of the runs, as previously seen in Equation 5.1. Due to the increased duration ofthe run, this reaction had more time to complete, thus removing the TeO2 from the system.Te and Na2Te were still present in the system, and the relative intensities of the peaks didnot appear to change significantly between the runs, implying no link between the relativeconcentrations and time. The recovery of tellurium to the slag phase was found to decreaseas the time increased, reducing from 97% at 15 minutes, to 87% at 75 minutes. This wasattributed to the volatilization of Te species, which were removed from the system in a gasor dust form. Since the furnace had its lid ajar, a constant flow of air into and out of thesystem was present, therefore allowing Te to leave the system. If Te was constantly beingremoved via the gas phase, then the recovery to the slag phase would in turn decrease.The silver recovery to the slag phase was found to decrease as the time increased. A1305.5. Processed Slag Analysis and Proposed Reactionsrecovery of 31.6% was observed for 15 minutes, compared to a recovery of 10.6% at 75minutes. XRD plots showed that an Ag2Te phase was present for all times examined,however the decreasing Ag recovery implies that there should be less of the Ag2Te present.Silver telluride (Ag2TeO3) was also seen in the 15 and 75 minute runs, while not appearingin the center point. This could again be due to the single XRD scan not showing the species.The presence of this species at shorter durations implies that the Ag2Te may react withother species (such as Na2TeO3) to form this silver telluride. These reactions are likely anintermediate reaction in the reduction of silver.5.5.3 Effect of PbO AdditionFigure 5.31 shows the XRD spectra comparing tests containing 0, 10, and 20 g leadoxide addition per 50 g slag. The species found to change between runs were tellurium,sodium telluride, tellurium dioxide, lead telluride, sodium tellurate, and silver telluride.Figure 5.31: XRD comparison of tests with 0, 10, and 20 g PbO addition per 50 g slag.Fixed inputs were 1000 ◦C pyrometallurgical processing temperature for 45 minutes, 5 g ofNa2CO3 / 50 g slag, and a pre-roasting temperature of 400◦C..When no PbO was added, the Pb recovery to the slag phase was found to be 13.2%,1315.5. Processed Slag Analysis and Proposed Reactionsas there was some intial Pb present in the slag. Upon the addition of 10 g of PbO, thePb recovery was found to decrease to 2.9%, likely due to the formation of a metal phasewhich absorbed any liquid metal formed. When the PbO addition increased to 20 g, thePb recovery to the slag phase was found to be 9%, due to the formation of PbTe in theslag phase. The PbO was expected to react with carbon at temperatures starting at 300◦C according to the following reaction:PbO + C = Pb + CO (5.7)It was likely that all of the PbO was not removed from the system according to theabove reaction. This unreacted PbO existed in the slag phase as a liquid (Tmelt = 888◦C).From the XRD plot, the addition of PbO into the system was found to produce PbTe inthe slag phase as the PbO addition increased. Subsequently, at this high addition amountof PbO, the Na2Te peak was seen to decrease in relative intensity compared to Te. It wasassumed that the Na2Te reacted with the Pb or PbO to form PbTe. Te may also react withPb or PbO to form PbTe. The equations for the reactions of both Te and Na2Te are:Pb + Te = PbTe (5.8)PbO + Te = PbTe + 1/2O2 (5.9)Pb + Na2Te = PbTe + 2Na (5.10)PbO + Na2Te = PbTe + Na2O (5.11)The silver recovery to the slag phase with no PbO addition was found to be 53.3%.This high recovery can be attributed to the diminished size of the metal phase, where aPbO addition would help to form a larger metal phase, with Pb acting as a collector for theliquid metals. When 10 g of PbO was added to the system, the Ag recovery dropped from53.3% to 26.2%. Further increasing the PbO addition to 20 g led to an increase in the Agrecovery, to 35.3%. The removal of Ag from the slag as the PbO increased from 0 to 10 gcan be explained by the following reactionsPbO + Ag2Te = 2Ag + PbTe + 1/2O2 (5.12)Pb + Ag2Te = 2Ag + PbTe (5.13)The Te recovery to the slag phase was found to increase slightly as the PbO additionincreased. With no PbO addition, a Te recovery of 93.4% was obtained, while at 20 g PbOaddition a Te recovery of 95.7% was obtained.1325.5. Processed Slag Analysis and Proposed Reactions5.5.4 Effect of Na2CO3 AdditionFigure 5.32 shows the XRD spectra comparing test containing 1, 5, and 9 g sodiumcarbonate addition per 50 g slag. The species found to change between runs were tellurium,sodium telluride, tellurium dioxide, and lead telluride.Figure 5.32: XRD comparison of tests with 1, 5, and 9 g Na2CO3 addition per 50 g slag.Fixed inputs were 1000 ◦C pyrometallurgical processing temperature for 45 minutes, 10 gof PbO / 50 g slag, and a pre-roasting temperature of 400 ◦C.The silver recovery to the slag phase was found to decrease as the amount of Na2CO3addition increased. At 1 g addition, the Ag recovery was 35.3% compared to 12.5% at 9 gaddition. The reactions are likely those seen in Equations 5.5 and 5.6.The lead recovery to the slag phase was found to decrease as the Na2CO3 increased from1 to 5 g, with no change between 5 and 9 g addition. A Pb recovery of 9% was found for1 g addition, and a recovery of 2.9% at 5 and 9 g addition. From the XRD plot, the PbTephase seen in the 1 g Na2CO3 addition was not present at higher concentrations, likely dueto the following reactions:PbTe + Na2O = Na2Te + PbO (5.14)1335.5. Processed Slag Analysis and Proposed ReactionsPbTe + Na2O = Na2Te + Pb + 1/2O2 (5.15)From the XRD plots, it can be seen that at both high and low additions of Na2CO3,the Na2Te relative intensity was higher than that of Te. At 9 g of Na2CO3, there wasno Te detected in the XRD plots. The higher amount of Na2Te was not expected, sincethere should not have been sufficient Na for the formation of Na2Te. Equation 5.2 showsa possible reaction between TeO2 and Na2O, however at higher additions of Na2CO3 thereactions between Te and Na species to form Na2Te are:Te + 2Na = Na2Te (5.16)Te + Na2O = Na2Te + 1/2O2 (5.17)5.5.5 Comparison to ThermodynamicsFigure 5.33 shows the Ag as Ag2Te from the thermodynamic model. It can be seen thatthe Ag2Te decreased as temperature and Na2CO3 increase. From the XRD spectra, thetemperature and PbO plots were the only spectra seen with Ag2Te present, however as thetemperature and PbO addition increased, the Ag2Te peaks diminished or were not present,indicating a good fit with the thermodynamics.1345.5. Processed Slag Analysis and Proposed ReactionsFigure 5.33: Thermodynamic model results showing Ag as Ag2Te as a function of temper-ature, PbO, and Na2CO3 additions.Figure 5.34 shows the Te as Na2Te from the thermodynamic model. It can be seen thatbesides PbO, an increase in temperature and Na2CO3 led to 100% of Te being present asNa2Te. Since Te was present in all XRD plots in some amount besides the 9 g addition ofNa2CO3, the model was therefore unable to accurately predict the tellurium distribution,or that the system had not yet reached the equilibrium state.1355.5. Processed Slag Analysis and Proposed ReactionsFigure 5.34: Thermodynamic model results showing Te as Na2Te as a function of temper-ature, PbO, and Na2CO3 additions.Figure 5.35 shows the Te as PbTe from the thermodynamic model. It can be seen thatthe increase in temperature and Na2CO3 led to a decrease in the amount of PbTe formed,which was seen in the Na2CO3 XRD plots, as the amount of Na2CO3 increased the PbTepeak was not longer seen. For PbO addition, the XRD showed the formation of PbTe,however this was not seen in the model. As the PbO increased in the model, the Te asPbTe only increased from 0.05% to 0.08%, implying that the model did not accuratelypredict PbTe fromation, or that the equilibrium state was not reached for the experimentalsystem.1365.5. Processed Slag Analysis and Proposed ReactionsFigure 5.35: Thermodynamic model results showing Te as PbTe as a function of tempera-ture, PbO, and Na2CO3 additions.Te and TeO2In the thermodynamic model, the primary Te species observed was Na2Te. Te wasassumed to report to the metal phase; however, the highest amount of Te seen was 0.15%.TeO2 was not observed in any quantity of 0.01% or greater during the thermodynamicanalysis. The potential reasons for this occurrence was that the thermodynamic modelreports the speciation at the processing temperature, while the XRD was conducted at roomtemperature. Therefore, as the samples cool down to room temperature, the formation ofTeO2 may occur, leading to the appearance of TeO2 in the XRD plots while not showingup in the thermodynamic model.5.5.6 Combined Leach Input SlagFigure 5.36 shows the XRD spectrum for the combined leach input material. Pulverizedpyrometallurgical products from tests with similar conditions were combined and mechan-ically mixed to produce a large volume of leaching input. It can be seen that there were1375.6. Pyrometallurgical Summarynine species which were able to be deconvoluted from the spectrum. However, due to thecomplex and partially amorphous nature of the slag, it is likely that several species werenot able to be identified which may have been important in the leaching process.Figure 5.36: XRD spectrum and species list for the combined leach input.5.6 Pyrometallurgical SummaryIn summary, from the 37 runs conducted, a statistical model was generated to determinethe factors which had the largest influence on the system. It was found that the statisticalmodelling did an adequate job at predicting the recovery and grade of the metal phasecompared to the slag phase. Rather than predicting the Te recovery to the slag phase, thismodel can be used to fine tune the recovery of certain elements to the metal phase, andthe rejection of others, such as Te. Since the Te had to report to either the slag or metalphases, recovery of the slag and a dust or gas capture system will allow for tellurium recovery.The resulting statistical models can be used as objective functions for minimization, in anattempt to find the ideal set of inputs for a desired output. The functions can also becombined, to optimize multiple output variables at once.XRD analysis of slag spectra revealed key tellurium species changes as certain inputschanged, leading to proposed reactions as certain inputs changed. These plots were com-pared to thermodynamic analysis, which found that the model was accurate in predictingAg recoveries, however the Te prediction was lacking. This either means that the modelrequires further refinement for use, or that the system had not reached an equilibrium state.When comparing the actual results to the theoretical results, the actual results for both1385.6. Pyrometallurgical SummaryAg and Te recoveries to their respective phases were lower than that of the model, whichfurther implies that the model requires adjustment, or that equilibrium was not obtained.139Chapter 6Results and Discussion -HydrometallurgyThis chapter examines the results of the leaching of tellurium and other relevant speciesfrom the processed slag. The concentration of several species and their recoveries are pre-sented. Results from the sodium sulphide cleaning of the solution are also shown.6.1 Preliminary Leaching ResultsResults from preliminary leaching tests are examined in this section. Figure 6.1 showsthe tellurium concentration in the solution as a function of slurry solids percentage. Itcan be seen that there was a linear correlation between concentration and slurry solidspercentage until 5% solids. Above this value, tellurium dioxide precipitation was observedduring ICP analysis, where an acidic internal standard was mixed with the basic sample,lowering the pH sufficiently to remove some tellurium from solution. From this, it wasassumed that the linear trend potentially continued until 10 or 12.5 % solids. At 15%solids, the tellurium concentration significantly deviated from the linear trend, implyingthat the leaching reactions had not reached completion or a similar equilibrium state to thelower solids percentage runs.1406.1. Preliminary Leaching ResultsFigure 6.1: Tellurium concentration as a function of slurry solids percentage in preliminaryanalysis.Figure 6.2 shows the effect of a sodium hydroxide addition and roasting on the leachingof tellurium from the processed slag. It can be seen that a 200 ◦C roasting step, with noNaOH addition, increased the Te recovery during leaching from 40 to 71%. Adding sodiumhydroxide to the roasting step resulted in further increase of the Te recovery. The proposedmechanisms for these phenomena were the oxidation of sodium telluride species into solublesodium tellurite, or reactions between sodium hydroxide, either during the roast or thesubsequent leach, where insoluble tellurium species were converted to soluble species.1416.2. Main Series Leaching ResultsFigure 6.2: Tellurium leach recovery as a function of sodium hydroxide addition, in weightpercentage of the input material, for roasted and not roasted runs during preliminary anal-ysis. The solids percentage for each test was 1% solids in slurry.6.2 Main Series Leaching ResultsThe results of the main leaching tests are examined in this section. For the majorityof the tests, two runs were conducted, with the line shown being the average of the twodata point collected. For the concentration bar plots, which show the final concentrationof a given element at the end of the test, the bar heigh represents the average of the twodata points collected, which are each visible on the plot indicated by a ‘+’ marker. In theconcentration plots, the red bar signifies a baseline run to which others were compared to.Certain tests were only able to conduct one run, so care was taken when making state-ments regarding the information presented by that data. While the data for the single runsmay be an outlier or contain other errors, they were included in the interest of discussionand for the potential for future tests.Results for the leaching behaviour of aluminum, arsenic, selenium, and silicon can befound in Appendix E, along with a brief description of other minor elements such as silver1426.2. Main Series Leaching Resultsand copper.6.2.1 Effect of Pre-Leach RoastingFigure 6.3 shows a comparison of the XRD spectra from the processed slag B and theroasted processed slag B. While the spectra are similar in shape, it was observed that theratio of relative intensities between the tellurium and sodium tellurite / sodium telluridepeaks changed after the roasting occurred. This suggests that some of the tellurium mayhave converted into sodium tellurite and/or sodium telluride during the roasting process.Sodium tellurite was the water soluble species of tellurium present in the slag, so an increasein the concentration in the slag should lead to the recovery of more tellurium during theleaching process.Figure 6.3: XRD spectra for the leach input and roasted leach input.Figure 6.4 shows a comparison of BSE SEM Images of a) the leach input and b) roastedleach input. It can be seen that compared to the regular leach input, the roasted leachinput contained more of the bright plate-like species (A). This species was determined tobe sodium tellurite via EDS analysis. This increase in the amount of sodium telluritepresent on the surface of the particles agrees with the XRD comparison in Figure 6.3,which implied that a conversion of tellurium into sodium tellurite may have occurred. Theincrease in sodium tellurite on the surface should lead to an increase in the Te recoveryduring leaching. Figures 6.5 and 6.6 show additional particles in the roasted leach input,showing more of the sodium tellurite present on the surface.1436.2. Main Series Leaching Results(a) Leach input. (b) Roasted leach input.Figure 6.4: BSE images of a) leach input and b) roasted leach input.1446.2. Main Series Leaching ResultsFigure 6.5: BSE image of the roasted leach input.1456.2. Main Series Leaching ResultsFigure 6.6: Higher magnification BSE image of the roasted leach input, showing the struc-ture of the sodium tellurite.6.2.2 Solution pHFigure 6.7 shows the solution pH as a function of leaching time for the as-received slagsand the processed slag B. The starting pH of the solution was 6. For all the slags, thepH of the solution immediately became basic once the leaching commenced. In the caseof slag A, the pH immediately reached a maximum, at which point it steadily decreasedover the 12 hour recording duration. Slag B and the processed slag B were observed toreach a maximum pH after approximately 1.5 to 2 hours into the leaching test. In both theas-received and processed slag B tests, the pH decreased as a the run continued, howeverthe rate of decrease was less than in the case of slag A. In all cases, the pH of the solutionat the end was less than 0.5 pH lower than the maximum value observed.1466.2. Main Series Leaching ResultsFigure 6.7: Solution pH as a function of time for as-received slag and processed slag Bleaching. Fixed inputs were the solids percentage at 5%, solution temperature of 25 ◦C,with no additives.The pH behaviour observed for the as-received slags indicates that the leaching of basicspecies occurred instantaneously in slag A, while in the case of slag B following the initialincrease to pH values of between 12 and 12.5, a slower release of basic species into solutionwould occur, leading to the delay in time when the maximum pH was observed. This couldbe caused by previously insoluble species able to be dissolved once the pH had reached athreshold value, leading to the release or exposure of basic species. In either case, the pHdecreasing as the leaching tests progresses could be attributed to the removal of hydroxideions from the solution, as seen in Section 6.2.3, where the tellurium recovery was found toincrease over the duration of all of these runs; thus, coupled with the pH decrease it waslikely that hydroxide species were reacting with insoluble Te species allowing them to bereleased into solution. Since each pH probe was calibrated prior to each run, the drift ofthe instrument measurement was not expected to contribute to the pH decrease.Figure 6.8 shows the pH as a function of time for the various slurry solids percentages.It can be seen that the pH of the solution, in most of the runs, increased as the solidspercentage increased. In all cases, the maximum pH observed occurred partway throughthe run, with the exception of the 15% solids run, where the pH continually increasedthroughout the run. This maximum pH value was associated with the point when all thebasic species had leached into solution. The pH would then decrease from the maximum1476.2. Main Series Leaching Resultsvalue as the run continued (with the exception of the 15% run).Figure 6.8: Solution pH as a function of time for various slurry solids percentages. Fixedinputs was solution temperature of 25 ◦C with no additives.Figure 6.9 shows the pH of solution as a function of time for initial NaOH additions of40 and 80 g/L. It can be seen that the pH of the solution when NaOH was initially addedwas over 13.7. For each test, it can be seen that the pH initially increased, followed by areduction in pH to values below the initial pH. This again implies that hydroxide ions wereleaving the solution, likely through reactions with solids in the residue.1486.2. Main Series Leaching ResultsFigure 6.9: Solution pH as a function of time for initial NaOH concentration. Fixed inputswere the solids percentage at 5% with a solution temperature of 25 ◦C.Figure 6.10 shows the pH of the solution as a function of time for the various solutiontemperatures. The results indicate that the increase in the leaching temperature led toa decrease in the pH of the solution. Interestingly, when the solution temperature wasraised from 50 to 60 ◦C the pH behaviour changed, where a maximum value was quicklyreached; afterwards the pH decreased throughout the remainder of the recorded data. Thiswas likely due to the increased temperatures leading to increased kinetics throughout thesystem, increasing the rate of leaching of hydroxide species into the system (hence theinstantaneous jump in pH), and increasing the rate of hydroxide species reacting withsolids, thus lowering the pH.1496.2. Main Series Leaching ResultsFigure 6.10: Solution pH as a function of time for various leaching solution temperatures.Fixed inputs were the solids percentage at 5% with no additives.6.2.3 Tellurium Leaching BehaviourFigure 6.11 shows the tellurium recovery as a function of time for the two as-receivedslags and the processed slag B. It can be seen that slag A yielded the highest recoveryof tellurium, followed by the processed slag B, with slag B yielding the lowest recovery.The differences in the recovery in these cases can be attributed to a difference in the slagchemistry. Furthermore, the pyrometallurgical processing step had a beneficial effect onprocessed slag B, since the Te recovery was higher by approximately 8 to 12%. Interestingly,both slag B and its processed form both featured an initial rapid increase in Te recoveryfollowed by a gradual increase until the end of the run. Slag A saw this initial sharp increaseas well, but was followed by a more gradual increase until approximately two hours intothe experiment, at which point the steady increase seen in both slag B and processed Bwas recorded. This could potentially be attributed to the differences in the soluble speciesdistribution within the slags, where in slag B and processed B the readily leachable materialswere able to be directly leached as soon as the run was initiated. Slag A experienced aslower dissolution after the initial increase, implying that fresh leachable species were beingexposed over the first two hours of the run.1506.2. Main Series Leaching ResultsFigure 6.11: Tellurium recovery as a function of time for the as-received slags and theprocessed slag. Fixed inputs were the solids percentage at 5%, solution temperature of 25◦C, with no additives. Two runs were conducted for each test.Figure 6.12 shows the concentration of tellurium in solution after the leaching testconcluded after 24 hours. It can be seen that the solution concentration followed the resultsseen in the previous recovery plot (Figure 6.11), where the solution of slag A had the highestconcentration, followed by the processed slag B, with the as-received slag B having the lowestconcentration. The processed slag B bar is red as it was used as baseline comparison.1516.2. Main Series Leaching ResultsFigure 6.12: Tellurium solution concentration after 24 hours for the as-received slags andthe processed slag. Fixed inputs were the solids percentage at 5%, solution temperature of25 ◦C, with no additives. Two runs were conducted for each test.Figure 6.13 shows the tellurium recovery as a function of time for various slurry solidspercentages. For each test, the tellurium recovery increased during the first 15 minutes tobetween 23 and 28%. This initial increase in the recovery can be attributed to the dissolutionof sodium tellurite (Na2TeO3) [32]. For the remainder of each test, as the amount of solidsincreased, the recovery of tellurium decreased, dropping from 61.86% (1% solids, averagedbetween two points) to values between 25 and 31% for the 30 and 10% solids, respectively.The increase in Te recovery following the initial increase can be attributed to the dissolutionof the previously non-soluble Te species such as elemental Te or TeO2. These non-solublespecies likely reacted with free sodium or hydroxide ions in the solution to form solublespecies. The pH figures seen in the previous section show a decrease in the pH as the runsprogressed, implying the removal of hydroxide ions from solution.1526.2. Main Series Leaching ResultsFigure 6.13: Tellurium recovery as a function of time for various slurry solids percentages.Fixed inputs were the solution temperature of 25 ◦C, with no additives. Two runs wereconducted for each test.Figure 6.14 shows the tellurium concentration after 24 hours for various slurry solidspercentages. It can be seen that as the amount of solids increased, the tellurium contentof the solution increased as well. The averaged maximum concentration observed was1536.2. Main Series Leaching Results29.05 g/L Te, obtained during the 30% solids run. This high concentration implied that asolubility limit was not reached for runs at lower solids percentages. This further impliedthat although the leaching of tellurium can continue at lower solids percentages, it mayrequire additional reagents or time in order to recover more Te into solution.Figure 6.14: Tellurium solution concentration after 24 hours for various slurry solids per-centages. Fixed inputs were the solution temperature of 25 ◦C, with no additives. Tworuns were conducted for each test.Figure 6.15 shows the tellurium recovery as a function of time for 5 and 10% slurry solids.These runs were allowed to run for 96 hours, with samples taken every 24 hours. It wasobserved that the recovery of Te increased as more time elapsed. For the 5% slag run, therecovery steadily increased with time, with the rate of leaching appearing to increase in thelast 24 hours of the run. Conversely, the 10% run did not see an increase in recovery between24 and 48 hours; however, after 48 hours the recovery started to increase at a slightly lowerrate than observed for the 5% run. These tests imply that allowing the leach to continuebeyond the previously examined 24 hours allows for the recovery of more tellurium, likelydue to a kinetically slower leaching reaction involving previously insoluble tellurium speciessuch as elemental Te (Equation 2.26, Rhee et al. [29]) or sodium tellurate.1546.2. Main Series Leaching ResultsFigure 6.15: Tellurium recovery as a function of time for 5 and 10% slurry solids. Fixedinputs were the solution temperature of 25 ◦C, with no additives. Two runs were conductedfor each test.Figure 6.16 shows the concentration of tellurium in solution after 24 or 96 hours elapsedleaching time. Due to the higher Te recovery after 96 hours, the resulting solution was inturn higher in Te concentration.1556.2. Main Series Leaching ResultsFigure 6.16: Tellurium solution concentration for 24 and 96 hour leach runs for 5 and 10%slurry solids. Fixed inputs were the solution temperature of 25 ◦C, with no additives. Tworuns were conducted for each test.Figure 6.17 shows the tellurium recovery as a function of time for various initial sodiumhydroxide concentrations. It can be seen that as the initial amount of NaOH increased, thetellurium recovery also increased. The increase from 0 to 100 g/L resulted in <5% increasein recovery. This effect was less pronounced than lowering solids percentage in the slurry orincreasing the leaching time. This Te recovery increase supports the presumption that freehydroxide in solution reacts with other non-soluble species (such as elemental Te, Equation2.26, Rhee et al. [29]) to recover more tellurium.1566.2. Main Series Leaching ResultsFigure 6.17: Tellurium recovery as a function of time for various initial concentrations ofsodium hydroxide. Fixed inputs were the solids percentage at 5% and solution temperatureof 25 ◦C. Two runs were conducted for each test.Figure 6.18 shows the concentration of tellurium in solution after 24 hours for thevarious initial sodium hydroxide additions. It was observed that while the averaged Teconcentrations for the NaOH additions were higher, some of the individual points for the1576.2. Main Series Leaching ResultsNaOH runs overlapped each other, or in the case of the 40 g/L run, overlapped the 0 g/Lbaseline. Since the majority of the NaOH points were higher than the highest observed 0g/L run, it can be stated that the addition of NaOH improved Te solution concentration andconsequently the Te recovery; however, the effect of increasing the concentration cannot bedefinitively stated. Runs containing initial concentration of 20 and 60 g/L NaOH were seento have a steady increase in Te concentration without overlapping their data points. Suchan observation suggests higher initial NaOH concentrations further increasing Te recoveryduring leaching.Figure 6.18: Tellurium solution concentration after 24 hours for various initial concentra-tions of sodium hydroxide. Fixed inputs were the solids percentage at 5% and solutiontemperature of 25 ◦C. Two runs were conducted for each test.Figure 6.19 shows the recovery of tellurium as a function of time for various leachingtemperatures. An increase in the temperature led to a higher Te recovery; however, examin-ing Figure 6.20, after 24 hours of leaching the data points for the elevated temperature runsall overlap each other, while remaining above the baseline run. This implies that a highersolution temperature resulted in a higher tellurium recovery. These results are comparableto those seen in Guo et al., where researchers observed a slight increase in Te recovery whenincreasing the temperature from 30 to 90 [32]. Interestingly, the increase in temperatureled to a higher initial increase in the Te recovery, implying an increase in the amount of1586.2. Main Series Leaching Resultssodium tellurite leached at the start, or the leaching of previously non-reactive or kineticallyslower species due to the increased solution temperature. The higher temperature likely ledto changes in the leaching kinetics of some species, allowing for faster dissolution to revealfresh sodium tellurite for leaching.Figure 6.19: Tellurium recovery as a function of time for various leaching temperatures.Fixed inputs were the solids percentage at 5%, with no additives. Two runs were conductedfor each test.1596.2. Main Series Leaching ResultsFigure 6.20: Tellurium solution concentration after 24 hours for various leaching tempera-tures. Each slurry consisted of 5% solids. Fixed inputs were the solids percentage at 5%,with no additives. Two runs were conducted for each test.Figure 6.21 shows the recovery of tellurium as a function of time for various initialconcentrations of sodium sulphide. Since only one run of each sodium sulphide leaching testwas conducted, the following are only preliminary results. It was observed that the additionof Na2S appeared to have led to an increase in the Te recovery. This was expected, sincethe Na2S likely reacted with previously insoluble sodium tellurate according to Equation2.27 from Guo et al., as described in Section 2.3.2 [32]. As the initial concentration of Na2Sincreased, the initial tellurium recovery in the first hour appeared to substantially increase,while the tellurium concentration would increase gradually over the remainder of the run.Tests conducted with higher initial concentrations of Na2S appeared to have a higher rateof tellurium leaching throughout the run compared to those with lower concentrations.1606.2. Main Series Leaching ResultsFigure 6.21: Tellurium recovery as a function of time for various initial concentrations ofsodium sulphide. Fixed inputs were the solids percentage at 5% and a solution temperatureof 25 ◦C. One run was conducted for each test.Figure 6.22 shows the tellurium concentration in solution at the end of the leaching testfor the various initial concentrations of sodium sulphide. The final solution concentrationappeared to have increased as more Na2S was added.1616.2. Main Series Leaching ResultsFigure 6.22: Tellurium solution concentration after 24 hours for various initial concentra-tions of sodium sulphide. Fixed inputs were the solids percentage at 5% and a solutiontemperature of 25 ◦C. One run was conducted for each test.Figure 6.23 shows the recovery of tellurium as a function of time for roasted and notroasted slags. Only one test was performed for each of the roasted slags, so no definitivestatements can be made from this data. It can be seen that when the slag was roasted,the tellurium recovery for each of the slags appeared to have increased from its non-roastedvariant. In the case of the processed slag B and roasted slag B, the recovery more thandoubled from an averaged value of 30 to 67%. In the case of the as-received slag B, therecovery doubled from 22 to 46%, while for the as-received slag A, the increase was from43 to 63% after roasting. As described in Section 6.2.1, the roasting process appeared tohave converted some of the elemental tellurium into soluble sodium tellurite, a water solublespecies. This conversion explains the higher recovery of Te observed.1626.2. Main Series Leaching ResultsFigure 6.23: Tellurium recovery as a function of time for roasted and not roasted slags.Fixed inputs were the solids percentage at 5%, solution temperature of 25 ◦C, and noadditives. One run was conducted for each roasting test, with two runs conducted for eachunroasted slag counterpart.Figure 6.24 shows the tellurium concentration of the solution for each slag leached,comparing the non-roasted and roasted variants. As expected, due to the tellurium recoveryincreasing as seen in Figure 6.23, the solution concentration of Te was therefore higher.1636.2. Main Series Leaching ResultsFigure 6.24: Tellurium solution concentration after 24 hours for roasted and not roastedslags. Fixed inputs were the solids percentage at 5%, solution temperature of 25 ◦C, andno additives. One run was conducted for each roasting test, with two runs conducted foreach unroasted slag counterpart.6.2.4 Tellurium PLSUpon completion of the leaching tests, the resulting leach solutions, after having samplestaken for ICP and a fraction set aside, were combined to form the primary Te pregnant leachsolution which would be used for the electrowinning tests. The leaching tests conductedwith the addition of Na2S were not blended into the main Te PLS, instead a second PLScontaining the solutions from the Na2S leaching tests was produced. Table 6.1 shows thecomposition for each of these leaching solutions.1646.3. Leaching Residue AnalysisTable 6.1: PLS compositions.Conc. (ppm) Te PLS Te + Na2S PLSAl 423.06 279.6As 16.27 9.63Pb 0.152 0.677S 159.93 2241Sb 21.51 17.42Se 31.84 33.7Si 327.63 184.45Te 7,090 6,035Na 18,530 11,0006.3 Leaching Residue AnalysisXRD analysis was performed on some of the leach residues in an attempt to determinewhich species were leached for the various conditions. However, due to the amorphousnature of the slags, it was likely that some species which are important to the leachingprocess were not identified. The legends in each plot show the species which change betweeneach spectrum. A XRD spectrum of the leach input slag with species can be found inChapter 5, Figure 5.36.Figure 6.25 shows XRD spectra of the leach input and the leach residue from the 1%solids test. It can be seen that sodium tellurite, aluminum silicon oxide, and aluminumselenate hydrate were all present in the leach input, and were not present in the 1% solidsresidue spectrum. These species were assumed to be leached into solution. Sodium telluriteis a soluble species, which was expected to be the main Te species leached into solution.Aluminum silicon oxide and aluminum selenate hydrate contributed Al, Si, and Se intothe leaching solution. Lastly, it can be seen that in the 1% solids residue, peaks for silvertelluride (Ag2Te) were observed. It was likely that the species were present in the inputmaterial, but in small quantities. The Ag2Te peaks were only seen once other species wereremoved from the solids.1656.3. Leaching Residue AnalysisFigure 6.25: XRD spectra of the leach input and 1% solids leaching residue.Figure 6.26 shows XRD spectra of the leach input and the leach residue from the 15%solids test. It can be seen that no new species appeared in the solids residue spectra,whereas four species were no longer observed, which were either leached or converted toother species. Sodium tellurite, sodium telluride, aluminum silicon oxide, and aluminumselenate hydrate were likely leached, as was observed in the previous 1% solids residue.Figure 6.26: XRD spectra of the leach input and 15% solids leaching residue.Figure 6.27 shows XRD spectra of the leach input and the leach residue from the 60 g/LNaOH run. The peak differences were similar to those observed in the 1% solids run, where1666.3. Leaching Residue Analysissodium tellurite, aluminum silicon oxide, and aluminum selenate hydrate were no longerdetected after the leaching occurred. A silver telluride peak appeared, likely being presentin the leach input but masked by other peaks or in small quantities below the detectionlimit.Figure 6.27: XRD spectra of the leach input and 60 g/L NaOH addition leaching residue.Figure 6.28 shows the XRD spectra of the leach input and the leach residue from the20 g/L Na2S run. Similar to the previous XRD spectra, sodium tellurite, aluminum siliconoxide, and aluminum selenate hydrate were not seen in the 20 g/L Na2S residue. Sodiumtelluride was observed in the leach input, but the peaks were not seen in the 20 g/L Na2Sresidue. Due to the addition of Na2S, potential reactions between Na2S and Na2Te orNa2TeO4 to form Na2TeS4 may have occurred, which would account for the increased Terecovery in the runs with Na2S. Unlike the previous runs, no new peaks were observed inthe residue.1676.4. Leaching SummaryFigure 6.28: XRD spectra of the leach input and 20 g/L Na2S addition leaching residue.In summary, peaks for sodium tellurite, aluminum silicon oxide, and aluminum selenatehydrate were not observed in any of the residue samples. This implied that they wereleached or converted to other species. These three species accounted for the Te, Al, Si, andSe found in the solution. In the residue, certain runs were seen to contain silver telluride inthe residue, however this species was likely present in the leach input, but in quantities toosmall to observe. Lastly, sodium telluride peaks were observed in the leach input, but inthe 15% solids and 20 g/L Na2S residues, these peaks were no longer present due to beingleached into solution, or through being merged with the larger Te peak close to it, makingXRD spectrum deconvolution challenging.6.4 Leaching SummaryA summary of the leaching results for the main five elements examined and the effectof the leaching parameters are shown in Table 6.2.1686.4. Leaching SummaryTable 6.2: Leaching summary.Solids Fraction Time NaOH Temp. Na2S RoastingTe HH NN N N NN NNSe – N – – NN NNSi HH H NN* N NN HHAl HH NN NN N NN –As N N N N N NThrough XRD analysis, the primary Te species leached was sodium tellurite, as expected.In runs where the Te recovery was higher, other Te species such as elemental Te, TeO2, orNa2TeO4 may leach into solution through kinetically slow reactions which produce solubleTe species. The increase of the slurry solids was found to have a negative effect on theleaching of Te, where the recovery was found to decrease from 61.86 to 31.39% when thesolids amount was increased from 1 to 10% slurry solids. This effect was offset when the leachwas allowed to run for more than 24 hours, where a 96 hour run led to a 48% Te recoverywith a 10% slurry solids percentage. This implies that a higher slurry solids percentagemay be viable, however the leaching time required to obtain high Te recoveries significantlyincreases. This was similar to an effect observed by Rhee et al., where increasing the pulpdensity from 10 to 100 g/L slag required doubling the leaching time to obtain similar leachingefficiencies [29]. The mechanism behind this was likely related to kinetics, where reactionsbesides the initial leaching of sodium tellurite were slower when more Te was present insolution. Thus, at higher solids amounts, where there was more sodium tellurite to initiallydissolve, the increased amount of Te in solution led to these reactions (i.e. Equation 2.26)slowing down, leading to a lower Te recovery.When examining solution modifiers, the addition of Na2S was found to have a largeimpact on the Te recovery, where the addition of 40 g/L Na2S was found to double the Terecovery after 24 hours. This was due to the reaction between insoluble Te species and theNa2S to form soluble Na2TeS4. The addition of NaOH was found to have a minor effect onthe Te recovery, leading to an increase in recovery of approximately 5%. When the solutiontemperature was increased, the Te recovery was found to increase by approximately 8 to13%.Lastly, roasting the slag at 200 ◦C for 3 hours appeared to more than double the Terecovery. This was due to the conversion of insoluble Te species to soluble sodium tellurite1696.5. Te PLS Sodium Sulphide Cleaningduring the roasting process.6.5 Te PLS Sodium Sulphide CleaningSolution cleaning via Na2S addition was attempted on the final leach solution. Smallamount of Na2S was added to solution and stirred for 6 hours. ICP-OES samples of thesolution were taken at 2 hour increments. ICP-OES analysis of the solution did not showany change after the 6 hours of stirring, however, when the solution was filtered a slightdiscolouration in the filter paper was observed, potentially indicating the presence of aresidue. Filter paper samples from the 1 and 5 x Na2S runs were examined via SEM, wherethe presence of a precipitate was seen.Figure 6.29 shows BSE images of the filter paper from the 1x (a) and 5x (b) Na2Saddition runs. The residue seen was found throughout both of the samples of the filterpaper examined. It can be seen that the 5x Na2S sample contained more residue than the1x sample, as expected.(a) 1x Na2S Addn. (b) 5x Na2S Addn.Figure 6.29: BSE images of a) 1x and b) 5x Na2S addition filter papers.Figure 6.30 shows a BSE image of residue agglomeration on the filter paper from the1x Na2S addition run. Using EDS, the composition of the residue was determined to be14.17% Na, 14.56% Al, 14.6% Si, and 8.12% Te. Trace elements were 0.79% Pb, 0.58%Fe, and 0.26% Cu. The balance was oxygen. Each of the elements precipitated, with the1706.5. Te PLS Sodium Sulphide Cleaningexception of the Te, reduce the overall solution impurity, which was the objective of the Na2Saddition. If impurity elements can be precipitated out of solution prior to electrowinning,then a cleaner EW product may be obtained.Figure 6.30: Higher magnification BSE image of the 1x Na2S filter paper, showing anagglomeration of residue.Figure 6.31 shows a SE image of residue particles removed from the 5x Na2S filter paperand placed onto carbon tape. EDS analysis of the residue revealed a composition similarto that of the previously seen residue, 16.30% Na, 15.64% Al, 15.85% Si, 2.51% Te, 0.15%Fe, with the remainder oxygen.1716.5. Te PLS Sodium Sulphide CleaningFigure 6.31: SE image of residue particles from the 5x Na2S addition cleaning test.In summary, the Na2S cleaning was able to produce a precipitate which yielded a cleanerPLS. While the initial objective was the removal of species such as iron as sulphides, theprimary species precipitated was a sodium aluminosilicate. This species was not expectedto be precipitated, but the resulting solution contained less impurities than before, whichwas an overall improvement. Some tellurium was found to be precipitated as well, whilethis was not ideal it was found that the addition of more Na2S (1x to 5x) led to a lowergrade of Te in the precipitate, 8.12 vs. 2.51%. The amount of Te precipitated in each runmay have been the same (in terms of mass), however the precipitation of additional sodiumaluminosilicate acted to depress the Te grade. Overall, the addition of Na2S in any amountserved as a means to clean the solution, which should improve downstream processes.172Chapter 7Results and Discussion - TelluriumExtractionThis chapter focuses on the extraction of tellurium from the tellurium leach solution.Two methods of tellurium extraction were examined: 1) electrowinning for the productionof metallic tellurium, and 2) tellurium dioxide precipitation via acidification.7.1 Preliminary Electrowinning ResultsResults from preliminary electrowinning tests are presented in this section. The resultsof these tests were used to determine the conditions for the main electrowinning tests.Figure 7.1 shows the effect of the current density (CD) and time on the telluriumconcentration in solution during electrowinning. Each data set was analysed using theMATLAB® Curve Fitting toolbox, which was able to be fit the results to a three-parameterexponential decay equation, as seen in Equation 7.1. It can be seen that an increase in thecurrent density led to an increase in the rate of tellurium depletion from the solution. Fora current density of 50 A/m2, 12 hours were required to reach 38% of initial telluriumconcentration in the solution, whereas increasing the current density to 150 A/m2 achievedthe same Te solution concentration in only 6 hours. Tellurium recovery values (inverse ofconcentration data shown in Figure 7.1) after 6 hours were 44, 62, and 72% for 50, 150, and250 A/m2, respectively. The final recovery for the 50 A/m2 test after 12 hours was 62%.1737.1. Preliminary Electrowinning ResultsFigure 7.1: Effect of current density and time on tellurium in solution for the preliminaryelectrowinning tests. The duration was 6 hours with a solution temperature of 25 ◦C withno additives.f(t) = a+ b · e−c · t (7.1)Where f(t) was the concentration of tellurium in solution at time t, in hours, and a, b, andc were constants. Values for each constant and the R2 for each curve fit are summarized inTable 7.1.Table 7.1: Parameters and R2 value for curve fit used for the preliminary electrowinningtests.CD a b c R250 A/m2 0.3085 0.6822 0.1743 0.995150 A/m2 0.1676 0.8279 0.2175 0.9848250 A/m2 -0.04429 1.045 0.1943 0.9999Table 7.2 shows the tellurium grade of the products obtained via ICP-MS analysis. It canbe seen that for all current density levels, 2N Te purity was obtained. The major impurities1747.1. Preliminary Electrowinning Resultsin each tellurium sample were sodium (0.574%), iron (0.067%), selenium (0.019%), copper(0.014%), lead (0.012%), and arsenic (0.009%). These impurities came from the solution,either from co-deposition or through entrapment of solution within pores in the product.For samples processed at 150 and 250 A/m2, additional impurities from electrode corrosionsuch as chromium (0.064%) and nickel (0.017%) were present.Table 7.2: Tellurium grades obtained via ICP-MS for preliminary electrowinning tests.CD Te Grade (%)50 A/m2 99.26150 A/m2 99.30250 A/m2 98.95Figure 7.2 shows the current efficiency (CE), calculated using Equations 3.1 and 3.2, asa function of the tellurium recovery for each test conducted. It can be seen that for eachtest, the current efficiency decreased as the tellurium recovery increased. A larger decreasein the current efficiency was observed as the current density decreased. For the 50 A/m2test the current efficiency decreased by of 49%, while for the 250 A/m2 test, the currentdensity decreased by only 7.4%, at 2 hours into the testing. These results show a trade-offbetween the efficiency of the process and the tellurium recovery. Practically, this impliesthat larger electrical costs will be incurred to reach higher levels of tellurium recovery; andafter a point it may become uneconomical to extract additional tellurium for the addedelectrical cost.1757.2. Electrowinning Tellurium RecoveryFigure 7.2: Current efficiency as a function of recovery for preliminary electrowinning tests.From the preliminary testing, it was decided that the benchmark test for the maintesting would be a current density of 150 A/m2. For this test, both the grade and recoveryobtained after 6 hours were used as a baseline with which other factors could be compared.7.2 Electrowinning Tellurium RecoveryFigures 7.3 through 7.7 show the effect of the various input variables on the telluriumconcentration in the leaching solution. Constant values for each test, unless otherwise stated,were 150 A/m2 current density, 50 cm2 cathodic surface area, 5 cm electrode spacing, 500mL of leaching solution, 60 RPM stirring, and a solution temperature of approximately 25◦C.For each test, the results obtained via ICP-AES analysis were plotted as a function oftime. The MATLAB® Curve Fitting Toolbox was used to fit to the 3-parameter exponentialdecay equation previously seen in Equation 7.1. The values of the constants a, b, and c foreach test can be seen in Table 7.3 below.1767.2. Electrowinning Tellurium RecoveryTable 7.3: Parameters and R2 value for curve fit used.Test a b c R250 A/m2 Run 1 -0.4307 1.434 0.06607 0.999450 A/m2 Run 2 -0.1503 1.159 0.08827 0.9939150 A/m2 0.1122 0.8887 0.2528 0.9992250 A/m2 0.1036 0.903 0.3507 0.9944Na2S PLS 0.05077 0.951 0.3443 0.997540 ◦C 0.04218 0.9623 0.3086 0.997950 ◦C 0.03439 0.9732 0.3563 0.97920.05 g/L Sucrose Addn 0.08452 0.9227 0.2501 0.99640.1 g/L Sucrose Addn 0.09639 0.9039 0.2706 0.99981x Na2S 0.101 0.9021 0.2841 0.99862.5x Na2S 0.09914 0.9074 0.2763 0.99665x Na2S 0.1216 0.8835 0.2785 0.9974Figure 7.3 shows the effect of the current density on the tellurium concentration insolution as a function of time. It can be seen that all four data sets fit well to the exponentialdecay function, however the two sets at a current density of 50 A/m2 show a different shapecompared to the 150 and 250 A/m2 sets. This was due to the linearity of the 50 A/m2 data,which in fact fit well to a linear model of y = m ·x+ b, with R2 values of 0.9789 and 0.9582for the first and second runs, respectively. For comparison, when the 150 A/m2 data wasfit to a linear model, the R2 value was only 0.7557. The linearity of the 50 A/m2 tests wasdue to the differences in electrowinning behaviour. The 50 A/m2 tests exhibited a period ofstable Te growth, whereas the 150 and 250 A/m2 tests did not see this due to the unstablegrowth of the crystals and excessive hydrogen evolution which led to the deposited materialfalling off of the electrode [37]. This effect was observed in all electrowinning runs, eventhe two 50 A/m2 runs between 6 and 12 hours. The removal of the plated material allowedfor the deposition of new tellurium from the solution, which continued the process. It canbe seen that increasing the current density led to a faster rate of tellurium removal fromsolution during the first 12 hours of electrowinning. After 12 hours, it can be seen that the150 and 250 A/m2 sets nearly converged at approximately 20 hours. After 20 hours bothcurves appeared to approach an asymptote, suggesting that the electrowinning had reached1777.2. Electrowinning Tellurium Recoveryan equilibrium.Figure 7.3: Effect of current density on tellurium concentration in solution as a function oftime. Solid markers are the experimental data, while the dashed lines are the curve fits.Fixed inputs were the solution temperature of 25 ◦C and no solution additives.1787.2. Electrowinning Tellurium RecoveryFigure 7.4 shows the effect of solution temperature on the tellurium concentration insolution as a function of time. It can be seen that at 40 ◦C, the tellurium removal ratewas fastest. A similar effect was seen at 50 ◦C after 4 hours. This implies that increasingthe temperature led to a faster rate of tellurium recovery, and as such this approach couldbe used to reduce the overall process time. Both the 25 and 40 ◦C data sets fit well tothe exponential curve fit, however the 50 ◦C set had the lowest fit quality of all 12 testsexamined in this section.Figure 7.4: Effect of temperature on tellurium concentration in solution as a function oftime. Solid markers are the experimental data while the dashed lines are the curve fits.Fixed inputs were current density at 150 A/m2 and no solution additives.Figure 7.5 shows the effect of Na2S addition on the tellurium concentration in solution asa function of time. It can be seen that there was little effect of the Na2S addition on the Teconcentration in solution over time. With the addition of 1 or 2.5 times the stoichiometricamount of Na2S (with respect to the impurity concentration), there was a minor decreasein the final Te concentration (approximately 2%).1797.2. Electrowinning Tellurium RecoveryFigure 7.5: Effect of Na2S addition for leach solution purification on the tellurium concen-tration in solution as a function of time. Solid markers are the experimental data whilethe dashed lines are the curve fits. Fixed inputs were current density at 150 A/m2 and asolution temperature of 25 ◦C.Figure 7.6 shows the effect of sucrose addition on the tellurium concentration of thesolution as a function of time. Similar to the Na2S addition, there was little effect of thesucrose on the Te concentration. While the Te concentration did decrease during the tests,the change was small ( 2.5%). The sucrose was added to act as an inhibitor to crystalgrowth, in an attempt to create more stable Te growth [59].1807.2. Electrowinning Tellurium RecoveryFigure 7.6: Effect of sucrose concentration on tellurium concentration in solution as afunction of time. Solid markers are the experimental data while the dashed lines are thecurve fits. Fixed inputs were current density at 150 A/m2 and a solution temperature of25 ◦C.Figure 7.7 shows the effect of a Na2S PLS solution as compared to the standard PLSsolution as a function of temperature. There was a notable change in the Te in solution whena PLS generated using Na2S was electrowon. It can be seen that the tellurium concentrationin solution throughout the duration of the test was approximately 10% lower in the Na2SPLS. This difference in concentration can be attributed to the solution chemistry, previouslyexamined in Chapter 6. The use of the Na2S in the leaching process led to an overallcleaner solution (Section 6.2.4, Table 6.1), which contained fewer impurities and had areduced sodium concentration. The reduction in other ions present in the solution wouldhave allowed for easier tellurium extraction, as there would be fewer interactions whichwould impede the movement of Te from solution to the electrode.1817.3. Current EfficiencyFigure 7.7: Comparison of sodium sulphide leach solution and regular leach solution on thetellurium concentration in solution as a function of time. Solid markers are the experimentaldata while the dashed lines are the curve fits. Fixed inputs were current density at 150 A/m2and a solution temperature of 25 ◦C.SummaryIn summary, the variables found to influence the tellurium concentration in solutionwere the current density, temperature, and Na2S PLS. The addition of small amounts ofNa2S and sucrose were not found to have a substantial impact on the Te concentration,however since these tests were not duplicated this lack of an effect cannot be stated withcertainty.7.3 Current EfficiencyFigures 7.8 through 7.12 show the current efficiency as a function of tellurium recovery.The equation used to determine the current efficiency can be found in Chapter 3, Section3.4.3. The values used in the equation were: M = 127.6 g/mol and z = 4. The electriccharge value, Q, varied depending on the current density of the run and was multiplied bythe time elapsed. Plotting these variables against each other allows for a comparison ofprocess efficiencies, essentially how much tellurium can be extracted per amount of energyput into the system. In some cases, optimizing the process to be energy efficient may be1827.3. Current Efficiencydesired over Te recovery. The Te recovery can potentially be increased with additionalelectrowinning or precipitation steps, which may be preferred to a single step with highenergy cost.Figure 7.8 shows the current efficiency as a function of current density and telluriumrecovery. It can be seen that a decrease in current efficiency occurred when both the currentdensity and tellurium recovery increased. The current density had a greater impact on thecurrent efficiency than the tellurium grade. At current densities of 150 and 250 A/m2,the final Te recoveries were within 4% of each other, while the current efficiency was 10%lower at 250 A/m2. When the 150 A/m2 run was compared to the 50 A/m2 runs, the Terecovery was found to be 9 to 14% lower; however, the current efficiency was 54 to 60%lower compared to the 50 A/m2 runs.Figure 7.8: Current efficiency as a function of tellurium recovery and current density.1837.3. Current EfficiencyFigure 7.9 shows the current efficiency as a function of temperature and the telluriumrecovery. In general, as the temperature increased, the tellurium recovery increased forsimilar current density. The exceptions to this trend occurred during the 50 ◦C test, wherethe initial Te recovery and current density was the same as the 25 ◦C test. This trendcould be attributed to non-tellurium elements (impurities) depositing on the electrode priorto tellurium. Broderick et al. noted that lower impurity concentration could be achievedduring Te electrowinning by using temperatures lower than 45 ◦C [36]. This trend thenchanges with time, as for similar current densities the Te recovery essentially doubled. Asthe run continued, the recovery continued to increase to above 90%, while the currentdensity decreased, due to the tellurium deposition rate decreasing. Finally, at the end ofthe 50 ◦C run, the recovery increased by only 2% while the current density decreased by28% (comparing last to second last data point, i.e. 6 vs 12 hours). This ending point wassimilar to that of the 40 ◦C test. This could imply that an equilibrium had been reached,where no more tellurium can be removed from the solution.Figure 7.9: Current efficiency as a function of tellurium recovery and temperature.1847.3. Current EfficiencyFigure 7.10 shows the current efficiency as a function of Na2S addition and telluriumrecovery. It can be seen that the addition of Na2S typically led to a minor increase in thetellurium grade and current densities, however the increase in the current density decreasedas the tellurium recovery increased. At minimum current density values observed (∼27%),the addition of 1 and 2.5x Na2S led to a minor increase in the recovery.Figure 7.10: Current efficiency as a function of tellurium recovery and sodium sulphidecleaning.1857.3. Current EfficiencyFigure 7.11 shows the current efficiency as a function of sucrose addition and telluriumrecovery. It can be seen that the addition of sucrose, with the exception of the first point,led to an increase in the tellurium recovery for similar current densities. This was similarto the effect of adding Na2S as examined previously. This effect could be attributed to thesucrose in solution changing the deposition kinetics of tellurium. The sucrose was added toact as a colloidal substance in the solution, acting as an inhibitor for crystal growth and asmoother, more stable growth [59].Figure 7.11: Current efficiency as a function of tellurium recovery and sucrose addition.1867.4. Tellurium ProductFigure 7.12 shows the current efficiency as a function of Na2S PLS and tellurium re-covery. It can be seen that using the Na2S PLS resulted in an increase in both currentefficiency and tellurium recovery. This increase can be attributed to differences in the so-lution chemistry allowing for enhanced leaching of tellurium. Due to the Na2S PLS havingfewer impurity species and sodium ions present, the tellurium in solution had an easierpath from the solution to the electrode. As such, more tellurium could deposit withoutinterference, which resulted in a higher current efficiency.Figure 7.12: Current efficiency as a function of tellurium recovery and sodium sulphideleaching.7.4 Tellurium ProductThe tellurium products generated via electrowinning were analysed using ICP-MS, XRD,and imaged using SEM.7.4.1 Te GradeTable 7.4 shows the grade of the final tellurium product for each of the 12 tests. Figure7.13 is a graphical representation of the data in Table 7.4. Since triplicate tests were not1877.4. Tellurium Productable to be performed, the standard deviation data shown was determined using the relativestandard deviation data provided from the ICP-MS analysis, calculated using Equation 7.2.However, two tests at 50 A/m2 were performed, and were within their standard deviations.For the 150 A/m2 data, the red bar in Figure 7.13, was used as a baseline for comparisonpurposes due to obtaining the highest grade and intermediate recovery during preliminarytesting as seen in Section 7.1.%RSD =s|X| · 100 (7.2)Where s is the standard deviation and X is the sample mean.Table 7.4: Tellurium grade and standard deviations for products of each electrowinningtest, obtained via ICP-MS.Test Conditions Te Grade (%) Std. Dev. (%)50 A/m2-1 98.88 0.0450 A/m2-2 98.86 0.02150 A/m2 98.78 0.03250 A/m2 98.69 0.02Na2S PLS 99.89 0.0140 ◦C 97.95 0.0250 ◦C 96.64 0.070.05 g/L Sucrose 97.89 0.050.1 g/L Sucrose 97.88 0.271x Na2S 98.58 0.012.5x Na2S 98.30 0.055x Na2S 97.90 0.021887.4. Tellurium ProductFigure 7.13: Tellurium product grade for each of the 12 electrowinning tests. The red barfor 150 A/m2 denotes the baseline used to compare tests. The error bars show the standarddeviation, calculated from the %RSD obtained from the ICP-MS analysis.Despite variances in the testing parameters, all of the electrowinning tests were ableto achieve tellurium grades of over 96.5%. However, certain tests were able to achievesignificantly higher grades which was generally desirable for further purification. In orderto examine the effects of the variables, the baseline run of 150 A/m2 at room temperature,with no additions, was used for comparison.The current density was found to have a minor effect on the grade of the telluriumproduced. Lowering the current density to 50 A/m2 was found to increase the grade ofthe tellurium to 98.87%, averaged between the two runs. Conversely, an increase in thecurrent density led to a lower Te grade, 98.69%. The difference in grade between each ofthe current density steps was 0.09%, decreasing as the current density increases. This effectof decreasing grade can be attributed to the increased deposition of non-tellurium elements,or through the entrainment of solution within rapidly growing Te crystals, resulting in alower grade (as described by Handle et al. [35]).1897.4. Tellurium ProductThe addition of Na2S either as a solution cleaning agent or a leaching aid led to dif-ferences in the electrowinning of the solution. It can be seen that the use of Na2S in theleaching stage, resulting in the Na2S PLS, and subsequent electrowinning led to the highestTe grade achieved, 99.89%. This was due to the cleaner solution (i.e. less non-Te species)produced by the Na2S when used in leaching, as previously seen in Table 6.1.7.4.2 Impurity AnalysisFigure 7.14 shows the grade of the impurity species for each test conducted. It canbe seen that for each test except for the Na2S PLS test, the dominant impurity specieswas sodium (90% of the total impurity was sodium for these runs). The second highestimpurity was aluminum, followed by selenium. Other impurity species present in moderateamounts were copper, arsenic, selenium, silver, antimony, and lead. Species detected atminor levels (<0.0005%) were titanium, vanadium, neodymium, chromium, tungsten, tin,cobalt, and bismuth. The Na2S PLS had higher levels of copper, selenium, and antimony,which contributed to the majority of the impurity present. These impurities had two po-tential sources; 1) the slag and 2) the electrode. Due to the slag being a complex systemcontaining numerous elements, it was likely that some of these species originated from there,in quantities too small to measure in earlier processes. Alternatively, species such as Cr, V,and Ti may have been present in small quantities in the electrodes (i.e. 304 stainless steelhas 18-20% Cr), which were leached into the solution during electrowinning, either throughelectrochemical interactions or due to the basic nature of the electrolyte.1907.4. Tellurium ProductFigure 7.14: Grade of impurity species present in each electrowinning product.Figure 7.15 shows a comparison of the 150 A/m2 test (standard leach solution) andthe Na2S PLS product impurity grades. It can be seen that the dominant impurities inthe standard solution, sodium and aluminum, were still present, but at levels two ordersof magnitude lower. The copper, selenium, and antimony levels are seen to be an order ofmagnitude higher when the Na2S PLS was used compared to the standard leach solution.Other moderate impurity elements such as arsenic, silver, and lead are all one order ofmagnitude lower in the Na2S PLS product compared to the standard leach solution.1917.4. Tellurium ProductFigure 7.15: Impurity grade comparison of standard leach solution and Na2S PLS.7.4.3 XRD AnalysisFigure 7.16 shows the XRD spectra for five electrowinning tellurium products. Theseproducts were analysed to determine the effect of one variable compared to baseline results.Through the XRD analysis, there were three species able to be identified. Tellurium in itselemental form was found in all electrowinning products. Sodium tellurate hydrate was theprimary impurity species found to occur within three (or potentially four) of the products.The last impurity species was the non-hydrated form of sodium tellurate, which was onlyobserved in the sample electrowon at 50 ◦C.1927.4. Tellurium ProductFigure 7.16: XRD spectra for electrowon tellurium products.Sodium tellurate hydrate and sodium tellurate are impurities within the product whichmay affect further purification of the Te product. As seen in the impurity analysis sectionearlier, the 1% sodium in the products was likely contained within this sodium telluratehydrate phase, implying that some of the tellurium present in the sample was bound in thetellurate as opposed to its elemental state. This difference in speciation may act as reducingthe overall Te grade of the products for certain purification methods, as grades of 2N purityelemental Te may be required, not elemental plus bound Te. This phase appeared in all ofthe tests with the exception of the Na2S PLS condition, and to a minor extent, the sucroseaddition test.1937.5. SEM ImagesIt can be seen that at the baseline 150 A/m2 test condition, tellurium and sodiumtellurate hydrate were present. In the Na2S PLS condition, 0.01% Na was detected, so anysodium tellurate hydrate formed may not be present in high enough quantities to be seen viaXRD. The Na2S cleaning test (5x Na2S addition) was seen to have sodium tellurate hydratepeaks at higher intensities than in the baseline run. This 5x Na2S product contained 2%sodium, double that of the baseline, likely all contained within the tellurate phase, leadingto a more intense peak. The small peaks in the Na2S cleaning run occurring around 2θ of35 were from the sodium tellurate hydrate phase. The sucrose addition (0.1 g/L) did notresult in the sodium tellurate hydrate peaks, however at approximately the same location(2θ of 12.5) there was a flattened peak, which could be the tellurate. The lack of a telluratepeak is interesting as the 0.1 g/L sucrose run contained 1.8% sodium, but XRD did notidentify any phase which contained sodium. Finally, the 50 ◦C run contained minor peaksof sodium tellurate hydrate, and small, but well defined peaks of sodium tellurate. Thiswas the only run examined which contained these sodium tellurate peaks. The 50 ◦C runcontained the highest sodium content of any run at 3.2%, likely all contained within thesodium tellurate. The formation of sodium tellurate appeared to only occur at elevatedtemperatures, which may led to a higher sodium content as the sodium tellurate crystalgrowth continued.7.5 SEM ImagesSEM images of the electrowon products were taken in an attempt to observe visualdifferences between the samples. Figures 7.17 to 7.27 were taken within the span of twodays, using identical imaging settings. Images were taken using the in-beam detector forbetter resolution at higher magnifications.Figure 7.17 shows an image of the tellurium product produced at a current density of50 A/m2. It can be seen that the tellurium formation was not an even coating, but insteadas a coarse, uncontrolled growth of tellurium. These crystals were found to be needle-like(typical size 2-10 µm), with several void spaces between larger ‘branches’ of crystals. It wasassumed that some impurities seen may be due to these voids, due to entrainment of theleaching solution and subsequent deposition of aqueous impurities.1947.5. SEM ImagesFigure 7.17: InBeam SE image of the tellurium product electrowon at 50 A/m2 for 12 hoursat room temperature with no solution additives.1957.5. SEM ImagesFigure 7.18 shows an image of the tellurium product produced at a current density of150 A/m2. The crystal geometry was similar as in the case of the 50 A/m2 sample, howeverit can be seen that the crystals were more densely packed (i.e. fewer dark voids seen, surfaceappeared relatively ‘smooth’ compared to 50 A/m2), and were not as long as the previouslyobserved crystals (typical size 1-2.5 µm).Figure 7.18: InBeam SE image of the tellurium product electrowon at 150 A/m2 for 12hours at room temperature with no solution additives.1967.5. SEM ImagesFigure 7.19 shows an image of the tellurium product produced at a current density of250 A/m2. The Te crystals in this product are smaller than those in the previously observedproducts from 50 and 150 A/m2 (typical size 0.5 to 2 µm).Figure 7.19: InBeam SE image of the tellurium product electrowon at 250 A/m2 for 12hours at room temperature with no solution additives.1977.5. SEM ImagesFigure 7.20 shows an image of the tellurium produced at a current density of 250 A/m2.The phase running diagonally through the centre of the image (A) is the impurity phasesodium tellurate hydrate. This phase was observed to form as layered, hexagonal crystals,and pure tellurium was seen to agglomerate at the edges and surface of these impuritycrystals. This agglomeration may have been due to either the impurity acting as a nucleationsite for the Te, or the agglomeration occurred during the drying process of the product.Figure 7.21 contains a closer view of this impurity, where the growth of Te crystals can beseen on the edge of the impurity (bottom right).Figure 7.20: InBeam SE image of the tellurium product electrowon at 250 A/m2 for 12hours at room temperature with no solution additives.1987.5. SEM ImagesFigure 7.21: InBeam SE image of the sodium tellurate hydrate impurity in the telluriumproduct produced at 250 A/m2 for 12 hours at room temperature with no solution additives.1997.5. SEM ImagesFigure 7.22 shows an image of the tellurium product electrowon at 150 A/m2 at 40 ◦C.The structure of the crystal growth appears similar to those seen previously; however, thestructure of the crystals appears to contain more voids compared to the 150 A/m2 sample(dark spaces), which could potentially explain higher impurity amounts as evaporation ofentrained solution would deposit impurity species in solution onto the tellurium.Figure 7.22: InBeam SE image of the tellurium product electrowon at 150 A/m2 for 12hours at 40 ◦C with no solution additives.2007.5. SEM ImagesFigure 7.23 shows an image of the tellurium product electrowon at 150 A/m2 at 50◦C. This lower magnification image shows the bulk structure of the electrowon tellurium,which was unique amongst any of the tests examined. The formation of the tellurium wasin larger, columnar shaped units, compared to a bulk mass seen in Figures 7.17 and 7.19.The difference in structure was likely due to the deposition kinetics of the solution, it wasobserved in Figure 7.4 that the behaviour of the tellurium concentration in solution wasdifferent than that of the other temperature tests. The initial lag in the Te concentrationfollowed by rapid deposition was likely to be responsible for the unique geometry of theproduct.Figure 7.23: InBeam SE image of the tellurium product electrowon at 150 A/m2 for 12hours at 50 ◦C with no solution additives.2017.5. SEM ImagesFigure 7.24 shows an image of one of the tellurium columns seen in the previous image.The columns these crystals built up did not contain as many voids as previously seen.Instead, these columns were found to contain larger amounts of impurity species (A), as isseen in Figure 7.25. Similar to the 250 A/m2 test, larger, hexagonal crystal of an impurityspecies were observed. These were assumed to be sodium tellurate, as this was the dominantimpurity species seen in the XRD analysis.Figure 7.24: InBeam SE image of the tellurium product electrowon at 150 A/m2 for 12hours at 50 ◦C with no solution additives.2027.5. SEM ImagesFigure 7.25: InBeam SE image of the tellurium product electrowon at 150 A/m2 for 12hours at 50 ◦C with no solution additives.2037.5. SEM ImagesFigure 7.26 shows the bulk structure of the tellurium product electrowon at 150 A/m2with and addition of 0.05 g/L sucrose. Overall, the structure of the product appeared thesame to that of the 150 A/m2 test without sucrose addition.Figure 7.26: InBeam SE image of the tellurium product electrowon at 150 A/m2 for 12hours at room temperature with 0.05 g/L sucrose addition.2047.6. Electrowinning Results SummaryFigure 7.27 shows an image of the bulk tellurium product produced at 150 A/m2 fromthe Na2S PLS. The structure was again similar to those produced from the non-Na2S PLSsolution, however no impurity phases were observed in the Na2S PLS product. The bulkstructure appeared homogeneous, with random, coarse crystal growth occurring throughout.Figure 7.27: InBeam SE image of the tellurium product electrowon at 150 A/m2 for 12hours at room temperature from the Na2S PLS.7.6 Electrowinning Results SummaryFigure 7.28 shows a summary of the electrowinning results. The current efficiency for themajority of tests were similar to each other, with the exceptions being the 250 and 50 A/m2tests. The 50 A/m2 tests had the highest current efficiency, but at the expense of tellurium2057.6. Electrowinning Results Summaryrecovery, while the opposite was true for the 250 A/m2 test. The addition of sucrose andNa2S into the solution led to a small increase in the tellurium recovery compared to thebaseline test. Increasing the addition of Na2S led to a decrease in both Te recovery andgrade, while increasing the sucrose led to a minor increase in grade and decrease in recovery.Increasing the solution temperature was found to increase the Te recovery, while reducingthe grade. The run with the highest grade and recovery was the Na2S PLS run, howeverthe current efficiency for this run was below average.Figure 7.28: Scatter plot showing electrowinning grade versus recovery, where the area ofthe point relates to the current efficiency.The ideal electrowinning parameters from the tests conducted would either be the 50A/m2 run with a solution at room temperature an no reagent additions, or the Na2S additionduring the leaching stage. The Na2S run was able to achieve the highest grade, at 99.89%,which may be sufficient enough of a grade for further refining via distillation or zone refiningwithout further processing. This high grade comes at the expense of current efficiency, whichwould require more power to obtain this amount of tellurium compared to the 50 A/m2 test.While the grade and recovery of the 50 A/m2 run were not as high as the Na2S PLS run,2067.7. Tellurium Dioxide Precipitationthe higher current efficiency could be desirable in an industrial setting. The lower recoverycould be alleviated by recycling the barren solution, or a two step recovery process.7.7 Tellurium Dioxide PrecipitationResults from the precipitation of tellurium dioxide from solution are examined in thissection. In an attempt to determine the effects of each acid, the concentration of theimpurity species in solution were examined. It was assumed that when the species leftthe solution, they would do so as a precipitate, therefore polluting the tellurium dioxideproduct.Figure 7.29 shows the pH as a function of acid addition for various 10 M acids. Asexpected, the pH of the solution decreased as the acid addition increased. The H2SO4 acidreduced the pH of the solution roughly twice as much compared to HCL and HNO3, due todouble the amount of hydrogen present comapred to HCl and HNO2. In order to accountfor this, a normalized plot of acid addition vs pH was produced and is seen in Figure 7.30.The pH path was essentially the same between each test when normalized, where smalldeviations can be accounted for through experimental errors in either the amount of acidor solution utilized, or minor differences in the molarity of the acids.Figure 7.29: pH of the tellurium solutions as a function of 10 M acid addition. Fixed inputswere 100 mL of leach solution at room temperature.2077.7. Tellurium Dioxide PrecipitationFigure 7.30: pH of the tellurium solutions as a function of normalized acid addition. Fixedinputs were 100 mL of leach solution at room temperature.Figure 7.31 shows the percentage of aluminum in solution as a function of acid addition.It can be seen that for each acid, the Al concentration initially decreased, implying precipita-tion of an aluminum oxide. When the precipitate produced at 5 mL was analysed via SEM-EDS, the aluminum bearing species was determined to be natrolite (Na2Al2Si3O10·2H2O).The aluminum in solution would then increase as more acid was added, however when suf-ficient quantities of both HCl and H2SO4 were added, 12.5 and 10 mL, respectively, thealuminum would again precipitate out of solution. Nitric acid was found to have the oppo-site effect, where at additions of 7.5 mL and greater, the amount of aluminum in solutionwas 85% or greater. At 15 mL HNO3, all of the aluminum returned to solution, likely dueto the stronger oxidizing nature of the HNO3 as opposed to the other acids utilized.2087.7. Tellurium Dioxide PrecipitationFigure 7.31: Percentage of aluminum in solution as a function of 10 M acid addition. Fixedinputs were 100 mL of leach solution at room temperature.Figure 7.32 shows the percentage of arsenic in solution as a function of acid addition.it can be seen that each acid had a similar effect on the As. The addition of small amountsof each acid led to a decrease in solution, where the arsenic was expected to report to theprecipitate. As each acid concentration increased, the arsenic gradually returns to solution,before completely returning at 15 mL addition of each acid. Interestingly, the sulphuricacid addition had the highest retention of the As for similar additions of the other acids.2097.7. Tellurium Dioxide PrecipitationFigure 7.32: Percentage of arsenic in solution as a function of 10 M acid addition. Fixedinputs were 100 mL of leach solution at room temperature.Figure 7.33 shows the selenium in solution as a function of acid addition. It can be seenthat the HCl and H2SO4 acids had a similar effect on the Se, where additions of 7.5 to 10mL led to 90% of the Se reporting to the precipitate phase. As the acid additions increased,the Se returned to solution. HNO3 also sees this effect, but to a limited degree, with only30% of the Se being removed from solution.2107.7. Tellurium Dioxide PrecipitationFigure 7.33: Percentage of selenium in solution as a function of 10 M acid addition. Fixedinputs were 100 mL of leach solution at room temperature.Figure 7.34 shows the percentage of antimony in solution as a function of acid addition.It can be seen that as the acid addition increased, the amount of antimony in solutiondecreased to 1.6%. This trend reverses when more acid was added, as 90% of the Sb wasrecovered to solution after 15 mL of each acid were added.2117.7. Tellurium Dioxide PrecipitationFigure 7.34: Percentage of antimony in solution as a function of 10 M acid addition. Fixedinputs were 100 mL of leach solution at room temperature.Figure 7.35 shows the percentage of tellurium in solution as a function of the acidaddition. The trend observed with the tellurium was similar to that observed with theantimony, except that the lowest concentration in solution was 6%. Similarly to the Sb, asthe acid addition increased, the tellurium returned to solution. At 15 mL acid addition,all of the tellurium had returned to solution. The precipitation and re-dissolution of Tein solution was expected as described in Chapter 4, Figure 4.34. The Sb and Te bothreached their minimum concentration in solution at the same acid addition, which led tothe tellurium dioxide product becoming contaminated with antimony trioxide (determinedvia SEM-EDS).2127.7. Tellurium Dioxide PrecipitationFigure 7.35: Percentage of tellurium in solution as a function of 10 M acid addition. Fixedinputs were 100 mL of leach solution at room temperature.Figure 7.36 shows the tellurium dioxide grade and tellurium recovery as a function ofnormalized acid addition. It can be seen that the grade of the TeO2 increased with acidaddition to a maximum of between 82 and 97% at 10 mL acid, after which it redissolves intosolution, in the case of the HCl run. The grade of TeO2 for HNO3 and H2SO4 additions wasunable to be determined after 10 mL due to an insufficient quantity of precipitate forming.The recovery of tellurium increased when the acid addition increased from 5 to 10 mL,before dropping to approximately 50% for all tests at 10 mL addition. The recovery furtherdecreased for the HCl test at 12.5 mL. Further increases to the acid addition reduce the Terecovery to 0% as Figure 7.35 shows 100% of Te returning to solution.2137.7. Tellurium Dioxide PrecipitationFigure 7.36: Tellurium dioxide product grade and tellurium recovery as a function of nor-malized acid addition Fixed inputs were 100 mL of leach solution at room temperature.2147.8. Tellurium Dioxide SEM Images7.8 Tellurium Dioxide SEM ImagesSEM-BSE imaging and EDS analysis was conducted on the tellurium dioxide samplesproduced during the nitric acid addition tests.Figure 7.37 shows a BSE image of the product produced from the addition of 5 mL ofHNO3. When analysed via EDS, the sample was found to contain 14.5% Na, 17.6% Al,20.2% Si, 0.17%K, 0.8% Te, with the remainder being oxygen. The closest mineralogicalmatch was found to be natrolite (Na2Al2Si3O10·2H2O).Figure 7.37: BSE image of precipitate produced using 5 mL of 10 M HNO3 into 100 mL ofTe PLS at room temperature.2157.8. Tellurium Dioxide SEM ImagesFigure 7.38 shows a 25,000 x magnification image of the surface of the natrolite pre-cipitate. The sufrace was found to have an uneven texture with no obvious bulk crystalstructure being observed, with little to no voids or other cavities occurring within the sam-ple.Figure 7.38: BSE image of natrolite surface formed through addition of 5 mL of 10 M HNO3into 100 mL of Te PLS at room temperature.Figure 7.39 shows an image of the precipitate where two visible phases were observed.The darker grey phase was the natrolite as previously seen, while the brighter phase wasfound via EDS to contain 7.71% Na, 17.10% Al, 6.82% Si, 21.84% Te, and the remainderoxygen. The texture of the phases appeared the same, implying that the Te likely formedTeO2 and adhered to the natrolite. Since the product was filtered through a filter paper toseparate the residue from solution, the potential for the tellurium oxide to adhere to the2167.8. Tellurium Dioxide SEM Imagesnatrolite during the filtering was also considered. However, this filtration would likely leadto a uniform distribution of Te across all grains of natrolite, and not concentrated on asmall amount of them, such as the particle seen in the image.Figure 7.39: BSE image of additional precipitate produced using 5 mL of 10 M HNO3 into100 mL of Te PLS at room temperature.2177.8. Tellurium Dioxide SEM ImagesFigure 7.40 shows an image of the precipitate produced when 7.5 mL of HNO3 wasadded to the solution. EDS analysis determined the composition to be 7.14% Na, 6.97% Al,5.04% Si, 44.49% Te, and the remainder oxygen. The appearance of the sample was similarto those previously seen, implying that the tellurium and natrolite may be co-depositingand agglomerating together. The uniform nature of the samples may also indicate that thetellurium and natrolite may form separate phases which combine when filtered.Figure 7.40: BSE image of precipitate produced using 7.5 mL of 10 M HNO3 into 100 mLof Te PLS at room temperature.2187.8. Tellurium Dioxide SEM ImagesFigure 7.41 shows the sample produced when 10 mL of HNO3 was added to the TePLS. The shape of the precipitate was much different in this run compared to others, asthe structure of the precipitate was either spherical or agglomerated rhombohedral crystals.The figure below shows the spherical crystals, and Figure 7.42 shows these precipitates inhigher magnification. It can be seen that the spheres are a collection of plate-like crystals,arranged in the spherical manner. Sections of the sample can also be seen to contain therhombohedral structure. EDS analysis of the sample found the composition to be between72 and 79% Te, with the remainder oxygen. This implies that the product formed waspredominately TeO2.Figure 7.41: BSE image of spherical precipitate produced using 10 mL of 10 M HNO3 into100 mL of Te PLS at room temperature.2197.8. Tellurium Dioxide SEM ImagesFigure 7.42: Higher magnification BSE image of precipitates observed in Figure 7.41.2207.8. Tellurium Dioxide SEM ImagesFigure 7.43 shows the 10 mL sample, with a section of agglomerated crystals beingseen. Both these larger plates of crystals along with the spheres were observed. Figure 7.44shows a closer magnification image, 20,000 x, of the surface of one of these plates. Whenthe surface of these plates were analysed, the composition was found to be 80% Te and 20%O.Figure 7.43: Agglomerated precipitate observed using BSE detector.2217.8. Tellurium Dioxide SEM ImagesFigure 7.44: BSE image agglomerated precipitate surface as seen in Figure 7.43.2227.8. Tellurium Dioxide SEM ImagesFigure 7.45 shows an image of the filter paper used to separate the solids from solutionfor the 12.5 mL acid addition run. There was no visible precipitation in the solution,however when filtered a residue was seen under SEM. A higher magnification image can beseen in Figure 7.46. When analysed via EDS, the major elements were 13.5% Sb and 61.5%Te. The remainder was comprised of oxygen and carbon, which was from the filter paperbelow the sample.Figure 7.45: BSE image of filter paper after filtration of acidified Te PLS using 12.5 mL ofHNO3 into 100 mL of leach solution at room temperature.2237.9. Tellurium Dioxide Precipitation SummaryFigure 7.46: Higher magnification BSE image of filter paper after filtration.7.9 Tellurium Dioxide Precipitation SummaryIn summary, the precipitation of tellurium dioxide can be achieved through the additionof acid into the Te PLS. It was found that the type of acid did not have a substantial impacton the grade or recovery of tellurium, were adding a normalized 10 mL of 10 M acid led toTeO22 grades of between 82 and 87%, with a 50% Te recovery. The structure of the TeO2produced using these paramaters was found to be small, spherical particles comprised oflayers of randomly oriented plates, with larger agglomerations of rhombohedral crystals.224Chapter 8Results and Discussion - Flowsheet8.1 Process FlowsheetFigure 8.1 shows the flowsheet developed from the results of this research to achievethe extraction of tellurium from soda slag. For each step, the values which yielded thehighest recovery values during experimental testing are shown. Products and reagents areshown in rectangular boxes, processes are shown in boxes with rounded squares, and circlesrepresent a mixing step. For each step, the tellurium recovery is shown. While the recoveryof each step was moderately high, between 67 and 89%, when combined the overall processrecovery for tellurium was found to be 46%. The flowsheet, in the presented form, was alinear process, where no recycle loops in the stream were present. Silver recovery to themetal phase during the pyrometallurgical processing step was 92.11%, where the remainingsilver almost entirely reported to the residue in the subsequent leaching step. The Ag gradeof the Te product was low at 0.0013% Ag.Figure 8.1: Flowsheet for the processing of soda slag.2258.2. Suggested Flowsheet Improvements8.2 Suggested Flowsheet ImprovementsAs stated in the previous section, the flowsheet produced for the process was an openloop design. If recycle streams were implemented, it is likely that the recovery of both silverand tellurium could be increased. The following improvements could be made in an effortto improve the process:1) Pyrometallurgical Off-Gas Capture: The off-gas/dust from the pyrometallurgical processcontains the tellurium balance which does not report to either the slag or metalphase. In the above flowsheet this is approximately 11% of the Te. The capture andsubsequent processing of this dust/gas would be expected to increase the overall Terecovery of the process.2) Leach Residue Recycle: The leach residue in the flowsheet above contains approximately8% of the total silver and 28% of the total tellurium which entered the process.Recycling this residue back into the pyrometallurgical processing step could result inhigher recoveries of both Ag and Te.3) Barren Solution Recycle: The barren solution from the electrowinning process can berecycled by being used as the lixiviant for the leaching. The tellurium present inthis barren stream would be able to be electrowon again, increasing the recovery.A potential concern is the build-up of elements within this stream, such as sodium,which is leached into solution but not removed at any point. Since the soluble speciesis sodium tellurite, the potential could exist for a solubility limit of sodium to bereached.4) Barren Solution Processing: A second processing step on the barren solution may beconsidered as well, as a means to extract addition tellurium. This could be donethrough either a second electrowinning step, or through the precipitation of telluriumas tellurium dioxide through solution acidification.226Chapter 9Conclusion and Recommendationsfor Future Work9.1 Thermodynamic ModelA thermodynamic model was produced in HSC 9.1. 66 species from a potential can-didate list of 709 were selected based on their stability across the temperature range inthe model. Activity coefficients for silver, tellurium, and lead were obtained and adaptedfrom literature. The model aimed to determine the effects of temperature, pre-roasting,and additions of silicon dioxide, calcium oxide, sodium carbonate, and lead(II) oxide. Atemperature of 1016 ◦C was determined to yield the highest Ag/Te ratio and a Ag recoveryof over 99% to the metal phase. Te recovery to the slag phase at this temperature wasalso found to be over 99%. The addition of silicon dioxide was found to have a negativeeffect on the system, where the Ag recovery to the metal phase would decrease as moreSiO2 was added. Conversely, both sodium carbonate and calcium oxide were found to havea positive effect on the system, seeing increased recovery of Ag to the metal phase whilesimultaneously seeing a high Te recovery to the slag phase. Lastly, lead(II) oxide was foundto have a negative effect on the system, similar to that of SiO2. The formation of PbTe inthe metal phase as the amount of PbO addition increased led to an impure metal phase.When examining the effect of pre-roasting, the model showed that higher pre-roastingtemperatures led to better results after the high temperature processing. A roast temper-ature of 500 ◦C had the greatest effect, however the changes themselves were a relativelyminor deviation from the baseline results, for example the Ag/Te ratio of the metal phaseincreased by 28 from a baseline of over 1800.9.2 Pyrometallurgical TestingA summary of the results obtained from pyrometallurgical testing resulted in two pri-mary products; 1) a metal phase rich in Ag and Pb, and 2) a slag phase containing Te.2279.2. Pyrometallurgical Testing9.2.1 Central Composite Design ResultsA central composite design methodology containing 36 tests was developed using MATLAB®’sccdesign function. The results of the tests were then used as inputs for a full quadraticmodel using MATLAB®’s stepwiselm and fitlm functions. Four models were producedand are summarized below.Tellurium Recovery to the SlagThe model for tellurium recovery to the slag phase was not successful in statisticallyvalid prediction of experimental data. The model had a negative R2adj value, and a F -statistic of 0.888, meaning it was slightly better than a single (i.e. constant) term model.The p-value of the model was also found to be 0.529, which was the highest value seen.This high p-value meant that it was unlikely that any of the input variables had an effecton the response (Te recovery). Should the model results be taken as quantitatively correct,then the temperature and lead oxide additions had the largest effect on the recovery, wherean increase in both above the baseline values (1000 ◦C, 10 g PbO) led to an increase in thetellurium recovery. However, due to the low accuracy of the model, it should not be usedfor predicting results of the system.Tellurium Recovery to the MetalThe model for tellurium recovery to the metal phase had very good statistical results,with an R2adj value of 0.836, an F -statistic of 13.2, and a p-value below 1x10-5. Thesestatistical results imply that the model fit well to the data. However, negative valuesexisted in the model at some edge cases for input variables. As such, the model should beused with caution to avoid these areas, or reconstructed with some constraints to minimizeor remove the negative values. Temperature and time were found to be the dominant termsin the model, having the only two square terms the model had. Temeprature also featureda large linear term, and four interaction terms.Silver Recovery to the MetalThe model for silver recovery to the metal phase was found to have adequate statisticalsignificance (a R2adj value of 0.612, F -statistic of 8.1, and p-value of 0.0013). Temperatureand lead oxide addition were found to have the largest effect on the Ag recovery to the metal,with both of these input variables having square terms, which had the greatest effect, inthe model.2289.2. Pyrometallurgical TestingTotal Impurity Grade of the MetalThe model for total impurity grade of the metal phase was found to have the second beststatistical significance (R2adj value of 0.755, an F -statistic of 12.1, and a p-value of below1x10-5). Temperature was found to have the largest effect on the total impurity, wherea temperature region existed typically between 980 and 1060 ◦C with the lowest grades.Following temperature, the sodium carbonate addition and roast temperature were foundto have the second largest effects.CCD SummaryIn summary, certain statistical models could be used to predict the products of thepyrometallurgical processing. Since the models produced for the metal phase responsevariables (Te recovery, Ag recovery, total impurity grade) all had adequate statistical sig-nificance, prediction of the metal phase would be more reliable than predicting the slagphase results. As such, examining how to reject tellurium from the metal phase can beconsidered the preferred method for tellurium recovery, as the Te not present in the metalphase reported to either the slag or the gas phase, each of which can be collected for futureprocessing. The models developed can be used as objective functions for optimization ofthe system, where models could also be combined in various ratios as a means of increasingthe importance of one model over another.9.2.2 XRD ResultsX-ray diffraction analysis was conducted on the slag products of the star point runs (i.e.±2 runs), where in each set of runs examined only one variable changed between each runbetween two extreme points and the center point. This allowed for the determination ofthe effect of each variable on the system. Tellurium species seen in the input material wereTe, TeO2, and Na2TeO3, while the input for silver was Ag2Te.Effect of Temperature on the SlagAt temperature of 900 and 1000 ◦C, the primary tellurium species seen were elementalTe and Na2Te. As the temperature continued to increase to 1100◦C, the elemental Tewas replaced with Na2Te. Te recovery to the slag phase was found to reach a maximum,95%, at 1000 ◦C. At a lower temperature of 900 ◦C, the recovery was smaller at 89%, andat 1100 ◦C the recovery decreased to 88%. More tellurium reported to the metal phaseat lower temperatures, as the temperature increased the kinetics of the system increased,and allowed for the slag to clean the metal phase, removing Te from it. Finally, as the2299.2. Pyrometallurgical Testingtemperature continued to rise, tellurium species volatilized and were removed from thesystem as a dust or gas.Silver, in the form of Ag2Te, was present in the XRD profiles at 900 and 1000◦C, andwas absent at 1100 ◦C. The recovery of silver to the slag phase decreased as the temperatureincreased, from 71% at 900 ◦C to 9.5% at 1100 ◦C.Effect of Processing Time on the SlagAs the processing time increased, the recovery of tellurium to the slag phase decreased.A recovery of 97% Te was observed at 15 minutes, however at 75 minutes the recovery haddropped to 87%. This was attributed to the volatilization of tellurium species which wereremoved from the system via the constantly changing atmosphere. Tellurium dioxide wasfound to be present in the system at 15 and 45 minutes, but was absent at 75 minutes. Thisremoval was likely due to reactions between TeO2 and carbon, which would form elementaltellurium and carbon monoxide as a by-product.Silver recoveries to the slag phase were found to decrease with time, with a recovery of32% at 15 minutes compared to a recovery of 11% at 75 minutes. For all runs, Ag2Te waspresent in the XRD plots, implying (along with the recovery values) that the removal ofAg2Te from the slag phase was not completed.Effect of PbO Addition on the SlagThe primary effect with the increase in PbO addition was the formation of PbTe in theslag phase at 20 g PbO level. The Te recovery to the slag phase was relatively unaffectedby the increase in PbO, increasing from 0 g PbO to 20 g PbO (93.4% compared to 95.7%Te recovery). Silver recovery to the slag phase with no PbO addition was found to be 53%,this was attributed to the diminished size of the metal phase, where the addition of PbOwould add Pb to the system to act as a collector for metal phases and increase the size ofthe metal phase. As the amount of PbO added increased to 10 g, the Ag recovery to theslag phase decreased to 26%, and further additions of PbO reversed the trend, leading to ahigher recovery of Ag to the slag phase of 35%.Effect of Na2CO3 Addition on the SlagThe addition of sodium carbonate to the system reduced the Ag recovery to the slagphase and the speciation of tellurium at high addition of Na2CO3. At 1 g addition, the Agrecovery was found to be 35%, while at 9 g addition the recovery decreased to 13%. Fortellurium, at high additions of Na2CO3, the dominant form of Te shifted from elemental2309.3. Hydrometallurgical LeachingTe to Na2Te, which was expected as the additional sodium present in the system would beused to form this species.9.2.3 Comparison of Experimental Results to ThermodynamicsThe results of the CCD experiments were compared to the results of the thermodynamicmodel. Two responses were examined; 1) the recovery of silver to the metal phase, and 2)the recovery of tellurium to the slag phase. In each case, the experimental results werelower than those of the model. Silver recovery results were typically 10 to 50% lower thanpredicted, whereas tellurium recovery results were closer to that of the model, typically be-tween 5 and 20% lower than the model predictions. This showed that further improvementsto the process were possible to reach a state closer to that of equilibrium. Alternatively,the model itself may require further refinement to better predict the system.9.3 Hydrometallurgical LeachingLeaching tests were carried out on the as-received slag, as well as the processed slagproduced from the pyrometallurgical testing.9.3.1 Effect of Leaching Parameters on Te RecoverySeveral leaching parameters were examined, and the effect of each on the leaching oftellurium is summarized below. In each case, the pH of the solution was found to be verybasic, typically at values of 12.5 or higher. Experiments with NaOH addition or higher solidspercentages increased this value further, to above 13. During the run, the pH decreased asthe run progressed, which was likely due to the removal of hydroxide ions from solution viachemical reactions resulting in the formation of soluble species.Effect of Slag TypeThe two as-received slags and the processed slag B were leached over the course of 24hours. Slag A was found to have the highest Te recovery after the 24 hours at 42%, with theprocessed slag B at 30%, and the as-received slag B at 22% recovery. These tests showedthat the pyrometallurgical processing of the slag B improved the leaching characteristics ofthe slag, due to the increased concentrations of soluble sodium tellurite species.2319.3. Hydrometallurgical LeachingEffect of Solids PercentageThe effect of solids percentage was examined on the processed slag B. Slurries containingbetween 1 and 30% solids were examined. It was found that for each test, the recovery oftellurium rapidly increased to 25-30%. This rapid leaching was attributed to the leachingof the soluble sodium tellurite species. As each run progressed, the recovery was foundto increase when the solids percentage of the slurry decreased. The run having 1% solidsreached a recovery of 62%, decreasing to 32% for 10% solids, and 26% for 30% solids. Thedifference in recovery was attributed to the leaching of previously non-soluble Te speciesfrom the residue. The proposed reactions involved free sodium or hydroxide in solutionreacting with the non-soluble Te species, pulling them into solution as soluble species. Asthe concentration of solids increased, it was likely that a higher concentration of Te species insolution slowed down the kinetics of the leaching reaction of elemental tellurium (Equation2.26), reducing the Te recovery.Effect of TimeExperiments containing 5 and 10% solids were executed for 96 hours compared to theprevious 24 hours. It was found that for both 5 and 10% solids, the recovery of telluriumincreased as the leaching duration increased. The 5% solids saw an increase in recovery ofover 10%, while the change for the 10% solids was smaller ( 5% increase). These resultsshowed that durations of over 24 hours were able to leach additional tellurium into solution.This was attributed to a kinetically slow reaction leaching elemental Te into solution in thepresence of sodium hydroxide in solution. At longer durations, this reaction was able to becarried out further, leading to an increase in Te recovery to solution.Effect of NaOH AdditionSolutions with initial NaOH concentrations between 20 and 100 g/L were prepared, andused as the lixiviant with 5% solids. It was found that the addition of the NaOH led tohigher Te recoveries after 24 hours, however the change was small, only a 5% increase in therecovery was seen when comparing 0 g/L to 100 g/L NaOH. These tests however, showedthat the NaOH concentration of the solution played a role in the leaching of tellurium,where higher concentrations of sodium or hydroxide accelerated reactions with insoluble Tespecies to solubilize them.2329.3. Hydrometallurgical LeachingEffect of TemperatureThe solution temperature was varied between room temperature, 25 ◦C, and elevatedtemperatures of 40, 50, and 60 ◦C. It was seen that increasing the temperature from roomtemperature to 40 and 50 ◦C led to an increase in the tellurium recovery, from 30% at roomtemperature to 43% at 40 and 50 ◦C. Further increasing the temperature led to a decreasein the tellurium recovery, to 38%, compared to the lower temperatures.Effect of Na2S AdditionWhen sodium sulphide was added as a leaching reagent, the recovery of tellurium sub-stantially increased. Additions ranging between 5 and 40 g/L Na2S were examined. When40 g/L of Na2S was added, the Te recovery doubled compared to 0 g/L addition, from 30%to 63%. Small addition of 5 g/L Na2S also had a large effect on the system, increasing theTe recovery to 47%. This increase in recovery was attributed to reactions between Na2Sand insoluble tellurium species as shown by Guo et al. [32]. In each of these runs, the initialtellurium recovery saw a similar large increase to when no Na2S was added, and a constantincrease in Te concentration was observed over the duration of the run.Effect of RoastingThe effect of roasting on the slags prior to leaching was examined. Both the as-receivedslags and the processed slag B were roasted and leached. When compared to the non-roasted leach runs, the recovery of Te in each case was found to increase. For slag B,the Te recovery doubled from 22 to 45%, while the processed slag B saw an increase inrecovery from 30% non-roasted to 67.11% roasted. Two potential reasons for this increasewere proposed. First, physical deformation of the slag at elevated temperatures allowedcracks to form, increasing the surface area for leaching or exposing fresh sodium telluritewhich would normally not be exposed to the solution. Second, reactions during the roastingprocess involving tellurium, sodium, and oxygen leading to the formation of soluble speciesmay have occurred.9.3.2 Leach Solution CleaningOnce the leach solution was produced, tests involving the addition of Na2S to the solutionin an attempt to clean it were conducted. It was found that the addition of Na2S intothe solution would form a precipitate which could be filtered from the solution. Uponexamination, the precipitate was found to predominantly contain a sodium aluminosilicatespecies, with trace amounts of lead, iron, and copper. However, tellurium was also present2339.4. Tellurium Extractionwith a grade of approximately 8%. While the precipitation of tellurium was unwanted, theremoval of the other species led to an overall cleaner solution, which was assumed to bebeneficial in downstream processing.9.4 Tellurium ExtractionThe resulting mixed solution from the leaching tests, containing 7 g/L Te, was used asthe feed material for tellurium extraction tests. Two methods, electrowinning and telluriumdioxide precipitation, were examined.9.4.1 ElectrowinningElectrowinning of the tellurium leach solution was conducted on small batches of 500mL of solution, with 304 stainless steel electrodes. The tellurium product formed was agrey powder, consisting almost entirely of elemental tellurium. This product did not adhereto the electrodes well, and as such the solution required filtration to recover the Te aftereach run. SEM analysis of the product showed that the product was not neatly layered, butconsisted of a seemingly random arrangement of micro-scale hexagonal needles. The needlenetwork had voids between each needle, which could act as a place for impurities from thesolution to be left behind if any solution were to be entrained within the product.Effect of Current DensityIt was found that a higher current density led to a faster recovery rate of tellurium fromthe solution to the product. After 12 hours, 250 A/m2 current density led to only 10.5% ofinitial Te remaining in solution. However, the current efficiency after 12 hours for 250 A/m2was only 15%, compared to 75 to 80% for 50 A/m2. The 250 A/m2 run also had the lowestTe grade when comparing current density, at 98.69% Te. In contrast, while the 50 A/m2left 22% of tellurium in the solution, it had the highest grade obtained when comparing thecurrent densities (98.88%). As such, the lower current density was selected as the optimalrun to obtain the highest grade while maximizing current efficiency.Effect of TemperatureThe temperatures examined during the electrowinning were room temperature, 40, and50 ◦C. It was found that the tellurium recovery increased as the temperature increased,however after 12 hours the 40 and 50 ◦C runs obtained a similar recovery of 92%. The50 ◦C run was able to obtain a Te recovery of 90% after only 6 hours, implying that the2349.4. Tellurium Extractionkinetics were faster at higher temperatures. Temperatures of 40 and 50 ◦C had a currentefficiency 3% higher than room temperature, but recoveries almost 10% higher. Finally, thegrade of the product decreased as the temperature increased. At room temperature, thetellurium grade of the product was 98.78%, dropping to 97.95% at 40 ◦C, and 96.64% at 50◦C.Effect of Na2S AdditionThe addition of Na2S as a cleaning agent for the leach solution was not found to havea significant effect on the tellurium recovery. The current density also did not appear tobe affected significantly after 12 hours. In both cases, a minor increase in the telluriumrecovery, 2%, was observed when Na2S was added. However, the grade of product wasnegatively affected. When no Na2S was added, the Te grade was 98.78%, decreasing to97.90% as more Na2S was added.Effect of Sucrose AdditionThe addition of sucrose was not found to have a substantial impact on the Te recovery,increasing by 2.5% for 0.1 g/L sucrose addition. The current efficiency was seen to slightlyimprove for similar recoveries, due to the sucrose in solution affecting the deposition me-chanics of Te. The sucrose was added to act as a colloidal substance which inhibited themovement of Te, attempting to force a smoother deposition. Finally, the grade of the prod-uct decreased when sucrose was added, from 98.78% with no addition to 97.88% for both0.05 and 0.1 g/L sucrose.Effect of Leaching SolutionThe use of a Na2S leaching solution was found to have a substantial effect on thetellurium electrowinning. The recovery using the Na2S solution was 10% higher compared tothe standard leach solution. The current efficiency was found to be higher by approximately20% for similar Te recoveries. The grade was the highest obtained at 99.89% for the Na2Sleach solution. This addition improved both the leaching and electrowinning of tellurium,and should be considered for implementation into the process. However, this reagent wasnot selected due to it introducing large amounts of sulphur into the system, which may notbe desired in industrial-scale operations.2359.5. Future Work9.4.2 Tellurium Dioxide PrecipitationThe precipitation of tellurium dioxide was examined via the acidification of the solutionthrough the addition of nitric, hydrochloric, and sulphuric acids. It was found that thetype of acid used did not impact the results obtained. The addition of 10 mL of 10 M acidto 100 mL of leach solution was found to produce a tellurium dioxide product containing82 to 87% TeO2, with a tellurium recovery of 50%. Antimony was the dominant impurity.The resulting product consisted of small, spherical particles comprised of layers of randomlyoriented plates, with larger agglomerations of rhombohedral crystals. This process couldbe used as an alternative to the electrowinning of tellurium, or as a intermediate step if theleach solution contains impurities. This method can also be used as cleaning step to removethe Te in the barren electrowinning solution.9.5 Future WorkFrom the work done on this dissertation the author recommends the following for futuresteps to improve the process developed.First, recovery improvements for each step of the process. The current work serves asa starting point for the future optimization of each step. The recovery of tellurium fromthe slag provided was shown to be possible, however the low overall process recovery canbe improved by optimizing each step. Additional variables can be examined, such as otherreagents in the pyrometallurgical testing, or combining variables in the leaching tests, forexample lower solids percentages with NaOH or Na2S additions.Second, testing on additional slags. In the current work, Slag B was the focus due to itscomposition being the closest to that of historically produced slags. However, other slagsshould be examined as well, with the aim of producing a generalized process, which had itsinputs based on the composition of the slag (i.e. 5 g PbO addition per 20 g of Te in theslag).Third, improvement of the thermodynamic model. The model produced acted as ameans to determine the effect of certain reagents and conditions on the recovery of sil-ver and tellurium. However, the experimental results differ from that seen in the model.Improvements to the model are necessary in order to closer match the system.Lastly, optimization of the process flowsheet. In the current state, the process is anopen circuit. Adding closed loops to the flowsheet could act as a means to reduce reagentconsumption and/or increase Te recovery.236Bibliography[1] mindat.org. The mineralogy of tellurium, 2019. → pages 1, 4[2] Hoffman, King, Carapella, Oldfield, and Putman. Tellurium and Tellurium Com-pounds. 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Establishingthese pseudo-equilibrium boundaries required long periods of time, and as a result, forma-tion of the  phase is expected to be slow from the L + (Pb) phase. As a result, the phaseboundaries have uncertainties ranging from 1% at 200 ◦C to 4% at 0 ◦C [60].245A.1. Phase DiagramsFigure A.1: The bismuth-lead (Bi-Pb) system [60]. Reprinted with permission from SpringerNature Customer Service Centre GmbH: Springer Journal of Phase Equilibria. The Bi-Pb (Bismuth-Lead) System, N. A. Gokcen. Copyright 1992.A.1.2 Silver-Lead SystemFigure A.2 shows the phase diagram for the silver-lead (Ag-Pb) system. This calculateddiagram was validated by previously completed experimental work [61]. An importantfeature of this phase diagram is the miscibility gap seen at temperatures below 500 ◦C,implying that distinct phases form in the final product as the solid species will not form asolid solution over a range of compositions.246A.1. Phase DiagramsFigure A.2: The silver-lead (Ag-Pb) system [61]. Reprinted with permission from SpringerNature Customer Service Centre GmbH: Springer Bulletin of Alloy Phase Diagrams.The Ag-Pb (Silver-Lead) System, Karakaya and Thompson. Copyright 1987.A.1.3 Lead-Tellurium SystemFigure A.3 shows the phase diagram for the lead-tellurium (Pb-Te) system. The verticalline at approximately 50 atomic percent tellurium is a separate phase, PbTe, which splitsthe L + PbTe phase. A phase diagram with a closer view of this phase is shown in FigureA.4 [62].247A.1. Phase DiagramsFigure A.3: The lead-tellurium (Pb-Te) system [62]. Reprinted with permission fromSpringer Nature Customer Service Centre GmbH: Springer Bulletin of Alloy Phase Di-agrams. The Pb-Te (Lead-Tellurium) System, Lin, Hsieh, Sharma, and Chang. Copyright1989.248A.1. Phase DiagramsFigure A.4: Close up of the PbTe section of Pb-Te phase diagram [62]. Reprinted withpermission from Springer Nature Customer Service Centre GmbH: Springer Bulletin ofAlloy Phase Diagrams. The Pb-Te (Lead-Tellurium) System, Lin, Hsieh, Sharma, andChang. Copyright 1989.A.1.4 Sodium-Tellurium SystemFigure A.5 shows the phase diagram for the sodium-tellurium (Na-Te) system. Sev-eral sodium-tellurium phases can be observed at the vertical lines at 73% atomic percenttellurium and above.249A.1. Phase DiagramsFigure A.5: The sodium-tellurium (Na-Te) system [63]. Reprinted with permission fromSpringer Nature Customer Service Centre GmbH: Springer Journal of Phase Equilibria.The Na-Te (Sodium-Tellurium) System, Pelton and Petric. Copyright 1990.A.1.5 Silver-Tellurium SystemFigure A.6 shows the phase diagram for the silver-tellurium (Ag-Te) system. A sec-ond phase diagram, showing a closer view of various phase boundaries at atomic percenttellurium values between 30 and 40% [64].250A.1. Phase DiagramsFigure A.6: The silver-tellurium (Ag-Te) system [64]. Reprinted with permission fromSpringer Nature Customer Service Centre GmbH: Springer Journal of Phase Equilibria.The Ag-Te (Silver-Tellurium) System, Karakaya and Thompson. Copyright 1991.251A.1. Phase DiagramsFigure A.7: Section of the silver-tellurium (Ag-Te) system [64]. Reprinted with permis-sion from Springer Nature Customer Service Centre GmbH: Springer Journal of PhaseEquilibria. The Ag-Te (Silver-Tellurium) System, Karakaya and Thompson. Copyright1991.252Appendix BSchematic DrawingsB.1 Experimental Design DrawingsFigure B.1 shows a dimensioned drawing of the graphite crucibles used during pyromet-allurgical testing. Units are millimetres (mm). The values shown are measured from one ofthe crucibles used. The critical dimension is the diameter of the column leading from thebase of the crucible to the bottom of the cone. This diameter must be within +0.1 mm ofthe diameter of the dowel pin inserted into the hole, otherwise the molten metal may flowaround the pin and out of the crucible.Figure B.1: Dimensioned drawing of the graphite crucibles used during pyrometallurgicaltesting. Dimensions are in millimetres (mm).253B.1. Experimental Design DrawingsFigure B.2 shows a dimensioned drawing of the electrodes used for electrowinning test-ing. The electrodes were machines from 0.81 mm thickness 304 stainless steel sheet, usinga waterjet. The edges were sanded down and faces polished prior to testing.Figure B.2: Dimensioned drawing showing the electrode used during main series electrowin-ning testing. All dimensions in millimetres.Figure B.3 shows a dimensioned drawing of the electrode holder used in electrowinningtesting. The slots for the electrodes are dimensioned larger than they were, to account forthe width of the waterjet stream when cutting the part. Final widths of the slots wereapproximately 1 mm. The three holes were used to take samples of the solution and toinsert probes or thermometers.254B.1. Experimental Design DrawingsFigure B.3: Dimensioned drawing showing the electrode holder used during electrowinningtesting. All units are in millimetres.255Appendix CSupplemental ThermodynamicModelling ResultsC.1 Effect of TemperatureAdditional figures for Section 4.6 are shown here.Figure C.1 shows the silver grade of the metal phase as a function of temperature. Itcan be seen that increasing the temperature resulted in an increase in the silver grade overthe entire temperature range examined. The largest increase in grade occurred at lowertemperatures, as the temperature increases. This increase diminishes at 950 ◦C as certainimpurity species are removed from the metal phase, either into the gas or slag phases.Another increase in the Ag grade occurred at 1100 ◦C, as Pb moved from the metal toeither the gas or the slag phase.256C.1. Effect of TemperatureFigure C.1: Silver grade of the metal phase as a function of temperature. Fixed inputs were100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.257C.1. Effect of TemperatureFigure C.2 shows the tellurium grade of the metal as a function of temperature. Itcan be seen that the grade rapidly decreased to almost zero as the Te reported to oxide orsodium species present in the slag. At temperatures abover 1100 ◦C, the tellurium beganto form greater amounts of PbTe, thus increasing its grade in the metal.Figure C.2: Tellurium grade of the metal phase as a function of temperature. Fixed inputswere 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.258C.1. Effect of TemperatureFigure C.3 shows the percentage of sodium present in the silicate species as a functionof temperature. Since the sodium-tellurium species in the slag phase are the ideal form oftellurium for further processing, the development of the sodium silicate species may reducethe recovery of tellurium in downstream processing stages. It can be seen that the sodiumin silicate species increased to a maximum of 15% at 1000 ◦C, rapidly decreasing beyondthis temperature as the silicate species (e.g., *2Na2O*SiO2) form other, more stable silicatescontaining less sodium (e.g., Na2O*2SiO2).Figure C.3: Sodium present in silicate species as a function of temperature. Fixed inputswere 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.Figure C.4 shows the tellurium grade of the slag phase as a function of temperature.Similar to the Ag grade of the metal phase, the Te grade of the slag phase increased as thetemperature increased; however, at temperatures of 1100 ◦C and higher, the rate began todecrease. The increase in grade can be attributed to species such as lead and silver reducingto their elemental forms as temperature increases, leaving the slag phase and reporting tothe metal phase. The continued increase in grade at temperatures of over 1000 ◦C, whenAg and Pb have settled in the metal phase, can be attributed to the breakdown of sodiumsilicate species, where the sodium would either report to the metal phase as a liquid or259C.1. Effect of Temperaturevolatilize to the gas phase. The decrease in the Te grade over 1150 ◦C was due to theincrease in Ag reporting to the slag as Ag2O and due to increased tellurium volatilizationto the gas phase.Figure C.4: Tellurium grade of the slag as a function of temperature. Fixed inputs were100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.260C.2. Effect of Fluxing ReagentsC.2 Effect of Fluxing ReagentsAdditional figures for Section 4.7 are shown here.Effect of Silicon DioxideFigure C.5 shows the Te/Ag ratio of the slag as a function of temperature and SiO2addition. Similar to the Ag/Te ratio trends, increasing the temperature cuased an increasein the Te/Ag ratio until a maximum at 1024 ◦C was reached. The addition of SiO2 decreasedthis ratio, since the formation of sodium silicate species led to an increase in the Ag2Te inthe slag phase.Figure C.5: Te/Ag ratio of the slag phase as a function of temperature and SiO2 addition.Fixed inputs were 100 kg of slag, 20 kg of PbO, and 10 kg of Na2CO3.261C.2. Effect of Fluxing ReagentsEffect of Sodium CarbonateFigure C.6 shows the Te/Ag ratio of the slag phase as a function of temperature andsodium carbonate addition. Similar to the Ag/Te ratio of the metal phase seen in Figure4.10, increasing both the temperature and Na2CO3 addition caused an increase in theTe/Ag ratio as the tellurium reported to the slag phase and the silver was removed fromthe slag phase to the metal phase. The maximum Te/Ag ratio occurred at a slightly highertemperature than the Ag/Te ratio for the metal phase, at 1042 ◦C as opposed to 1034 ◦C.Figure C.6: Te/Ag ratio of the slag phase as a function of temperature and Na2CO3 addition.Fixed inputs were 100 kg of slag and 20 kg of PbO.262C.2. Effect of Fluxing ReagentsEffect of Lead(II) OxideFigure C.7 shows the Te/Ag ratio of the slag phase as a function of temperature andlead(II) oxide addition. The addition of PbO led to an increase in this ratio, as did increasingthe temperature. The addition of PbO led to a decrease of silver in the metal slag, increasingthe Te/Ag ratio.Figure C.7: Te/Ag ratio of the slag phase as a function of temperature and PbO addition.Fixed inputs were 100 kg of slag and 10 kg of Na2CO3.263C.2. Effect of Fluxing ReagentsEffect of Lead(II) Oxide and Sodium CarbonateFigure C.8 shows the Te/Ag ratio of the slag phase as a function of sodium carbonateand lead(II) oxide addition. The addition of both Na2CO3 and PbO led to an increasein the Te/Ag ratio, with the highest ratio being achieved when the maximum amounts ofboth species were added. This was due to the reactions between Ag2Te and both PbO andNa2CO3 to form tellurate / telluride species in the slag phase, leaving Ag to report to themetal phase, increasing the Te/Ag ratio in the slag.Figure C.8: Te/Ag ratio of the slag phase as a function of Na2CO3 and PbO addition.Fixed inputs were 100 kg of slag and a temperature of 1000 ◦C.264Appendix DSupplemental PyrometallurgicalTesting ResultsD.1 Pyrometallurgical Experimental ResultsTable D.1 shows the recovery of tellurium to the metal and slag phases, the recovery ofsilver to the metal phase, and the total impurity grade of the metal phase for each of thepyrometallurgical tests conducted. Mass balancing the tellurium shows that not all of itreported to the metal or slag phases. This missing tellurium mass was assumed to reportto the dust/gas phase and can be seen in the table.265D.1. Pyrometallurgical Experimental ResultsTable D.1: Experimentally obtained values for the recoveries of tellurium and silver, andthe total impurity grade, used as input to the CCD model.Test No. TeSlagRec TeMetalRec AgMetalRec ImpurityMetalGr TeGasRec1 77.69 0.86 71.3 2.36 21.452 87.74 0.51 47.1 3.28 11.753 86.93 0.57 58.91 2.65 12.504 83.32 0.5 72.82 1.43 16.175 80.79 0.1 89.62 1.15 19.116 80.13 5.45 75.71 5.5 14.427 85.61 0.35 38.64 2.7 14.048 89.1 0.68 53.89 3.26 10.229 82.34 0.17 77.58 1.25 17.5010 84.83 1.05 81.91 1.69 14.1211 88.27 1.81 74.46 2.52 9.9312 82.85 0.04 51.35 2.09 17.1213 88.82 0.16 81.04 1.18 11.0214 93.16 0.01 71.06 1.18 6.8315 90.44 0.1 82.34 1.26 9.4616 88.46 0.14 92.11 1.08 11.4017 81.79 0.24 81.53 1.35 17.9718 87.01 0.22 65.09 2.17 12.7719-1 84.56 0.25 53.24 1.58 15.2019-2 89.13 0.18 50.91 1.72 10.6920 95.51 0.31 54.95 1.42 4.1921 93.53 0.31 61.4 1.5 6.1622 85.7 0.26 102.75 1.08 14.0423 88.78 3.31 29 4.95 7.9124 88.16 0.06 91.26 1.14 11.7825 93.41 0.16 78.03 1.33 6.4326 97.6 1.6 68.37 2.76 0.7927 86.98 0.36 89.37 1.13 12.6528 93.41 0.02 46.72 2 6.5729 94.68 0.12 73.76 1.24 5.2030 84.45 0.44 64.73 1.58 15.1231 95.77 0.32 64.73 1.58 3.9132 93.42 0.51 87.06 1.85 6.0733 89.22 0.15 89.74 1.25 10.6334 83.78 0.07 93.53 1.06 16.1635 91.86 0.2 71.6 1.31 7.9436 94.28 0.1 90.55 1.18 5.62266Appendix EAdditional Leaching FiguresAdditional results and figures pertaining to hydrometallurgical leaching results areshown here.E.0.1 Selenium Leaching BehaviourThe leaching behaviour of selenium is examined in this section. Due to selenium beinga chalcogen element, the leaching behaviour was expected to be similar to that of Te. SinceSe could be a contaminant in the electrowinning step, the leaching profile was importantto examine. Water quality guidelines are also present in British Columbia for Se in watersources, which should also be considered during the leaching process [65]. All of the runsexamined in this section had Se content above that of the water quality guidelines foraquatic life (2 µg/L).Figure E.1 shows the selenium recovery as a function of time for the as-received slags (Aand B) and the processed slag B. It can be seen that similar to the tellurium, slag A had ahigher recovery compared to slag B. This changed in the case of the processed slag B, wherethe recovery of the processed slag was higher than slag A. Similar to the Te behaviour, theSe recovery had a sharp increase in the initial recovery, followed by a gradual increase untila steady recovery value was reached. Figure E.2 displays the concentration of Se in solutionafter 24 hours of leaching. It can be seen that although slag A had a lower Se recoverycompared to the processed slag B, the concentration of Se in solution was similar betweenthese two runs (slag A 19 mg/L vs processed slag B 22 mg/L).267Appendix E. Additional Leaching FiguresFigure E.1: Selenium recovery as a function of time for the as-received slags and the pro-cessed slag. Each slurry consisted of 5% solids.268Appendix E. Additional Leaching FiguresFigure E.2: Selenium solution concentration after 24 hours for the as-received slags and theprocessed slag. Each slurry consisted of 5% solids.Figure E.3 shows the recovery of selenium as a function of time for various slurry solidspercentages. It can be seen that there was no observed difference after 24 hours in therecovery of Se to the solution. It can therefore be stated that the percentage of Se insoluble species in the processed slag B was initially between 21 and 23%. Figure E.4 showsthe concentration of selenium in solution after 24 hours of leaching. It can be seen that forthe 30% solids test, the averaged solution concentration of selenium was 175 mg/L, implyingthat the tests with a lower solids percentage were not reaching a solubility limit, furtheraiding the statement that only 22% of the selenium in the processed slag B was initiallysoluble.269Appendix E. Additional Leaching FiguresFigure E.3: Selenium recovery as a function of time for various slurry solids percentages.270Appendix E. Additional Leaching FiguresFigure E.4: Selenium solution concentration after 24 hours for various slurry solids percent-ages.Figure E.5 shows the selenium concentration of the leach solution after 24 and 96 hoursfor 5 and 10% slurry solids. Similar to the tellurium in the previous section, the additional72 hours of leaching allowed for the dissolution of more selenium. This was likely due to theexposure of fresh, soluble Se species, or through chemical reactions between components inthe solution and insoluble Se species in the residue.271Appendix E. Additional Leaching FiguresFigure E.5: Selenium solution concentration for 24 and 96 hour leach runs for 5 and 10%slurry solids.Figure E.6 shows the selenium concentration in solution after 24 hours of leaching forvarious initial concentrations of sodium hydroxide. It can be seen that there was no signifi-cant effect on the Se concentration as the initial concentration of NaOH increased, implyingthat the addition of free NaOH did not have an impact on the leaching behaviour of sele-nium.272Appendix E. Additional Leaching FiguresFigure E.6: Selenium solution concentration after 24 hours for various initial concentrationsof sodium hydroxide. Each slurry consisted of 5% solids.Figure E.7 shows the selenium concentration in solution after 24 hours of leaching forvarious solution temperatures. It can be seen that there was a minor increase in the solu-tion concentration between 25 and 50 ◦C. This minor increase was not observed when thetemperature was further increased to 60 ◦C.273Appendix E. Additional Leaching FiguresFigure E.7: Selenium solution concentration after 24 hours for various leaching tempera-tures. Each slurry consisted of 5% solids.Figure E.8 shows the selenium concentration in the solution after 24 hours of leachingfor various initial sodium sulphide concentrations. Since only one test for each Na2S con-centration was conducted, no definitive results should be drawn from the data presented. Itcan be seen that as the initial concentration of Na2S increased, the Se solution concentrationappeared to increase. This was likely due to the reaction between Na2S and insoluble Sespecies, similar to Equation 2.27 proposed by Guo et al. for Te [32].274Appendix E. Additional Leaching FiguresFigure E.8: Selenium solution concentration after 24 hours for various initial concentrationsof sodium sulphide. Each slurry consisted of 5% solids.Figure E.9 shows the selenium concentration in solution after 24 hours of leaching forroasted and not roasted slags, both as-received slags A and B and the processed slag B.The roasting appears to have increased the leachability of the selenium, as the solutionconcentration was higher for each slag than the baseline. Although not directly observed, theroasting may have affected the selenium-bearing species similar to the Te species, convertingthem into a more soluble species. Se may also have been contained within the soluble Tespecies (i.e. as an impurity), and dissolved into solution alongside the Te, leading to anincrease in recovery.275Appendix E. Additional Leaching FiguresFigure E.9: Selenium solution concentration after 24 hours for roasted and not roastedslags. Each slurry consisted of 5% solids.E.0.2 Silicon Leaching BehaviourThe leaching behaviour of silicon is examined in this section. Recovery values for siliconmay be marginally increased due to possible reactions between aqueous sodium hydroxide276Appendix E. Additional Leaching Figuresand the silica in the Pyrex beaker (Pyrex comprised of 53.9% O and 37.7% Si by weight[66]) and in which the leaching test was conducted.Figure E.10 shows the silicon recovery as a function of time for the as-received slagsand the processed slag. It can be seen that for both the as-received slags, the recoveryvalues over time were identical, following the same profile. However, when examining theSi solution concentration in Figure E.11, it can be seen that the Si concentration for slag Bafter 24 hours was approximately 8 times higher than that of slag A. The processed slag Byielded a higher recovery than slag B, while also yielding a higher Si solution concentrationby approximately 20 mg/L.Figure E.10: Silicon recovery as a function of time for the as-received slags and the processedslag. Each slurry consisted of 5% solids.277Appendix E. Additional Leaching FiguresFigure E.11: Silicon solution concentration after 24 hours for the as-received slags and theprocessed slag. Each slurry consisted of 5% solids.Figure E.12 shows the silicon recovery as a function of time for various slurry solidspercentages. It can be seen that as the amount of solids increased, the recovery of Si tosolution decreased. Interestingly, for each run the Si recovery would reach maximum valuesbetween 1 and 4 hours into the run, after which the Si recovery would decrease.278Appendix E. Additional Leaching FiguresFigure E.12: Silicon recovery as a function of time for various slurry solids percentages.Figure E.13 shows the silicon concentration in solution after 24 hours of leaching forvarious slurry solids percentages. It can be seen that a maximum averaged value of 653mg/L Si was observed in the 30% solids run. This high value implies that a solubility limitwas not reached in the lower solids percent runs.279Appendix E. Additional Leaching FiguresFigure E.13: Silicon solution concentration after 24 hours for various slurry solids percent-ages.Figure E.14 shows the silicon concentration in solution after 24 and 96 hours for 5 and10% slurry solids. It can be seen that the 5% run obtained a lower concentration of Si after96 hours compared to the 24 hour run, likely a continuation of the trend observed previouslyin Figure E.12. The 10% solids run did not see a significant change in the Si concentrationbetween 24 and 96 hours.280Appendix E. Additional Leaching FiguresFigure E.14: Silicon solution concentration for 24 and 96 hour leach runs for 5 and 10%slurry solids.Figure E.15 shows the silicon recovery as a function of time for various initial sodiumhydroxide concentrations. It can be seen that once the initial NaOH was added to thesolution, the Si recovery exceeded 100%. This was due to the reaction between NaOH andthe Pyrex beaker, leaching silicon into solution, in the form of sodium silicates. As the initialconcentration of NaOH increased, the Si recovery increased as well. Figure E.16 shows theSi concentration after 24 hours of leaching for the various initial NaOH concentrations. Itcan be seen that a maximum concentration of approximately 1 g/L for the 60 and 80 g/LNaOH runs.281Appendix E. Additional Leaching FiguresFigure E.15: Silicon solution concentration after 24 hours for various initial concentrationsof sodium hydroxide. Each slurry consisted of 5% solids.282Appendix E. Additional Leaching FiguresFigure E.16: Silicon solution concentration after 24 hours for various initial concentrationsof sodium hydroxide. Each slurry consisted of 5% solids.Figure E.17 shows the solution concentration of silicon for various solution temperatures.It can be seen that as the temperature of the solution increased, the Si concentrationincreased as well. This was likely due to faster leaching kinetics due to higher solutiontemperatures.283Appendix E. Additional Leaching FiguresFigure E.17: Silicon solution concentration after 24 hours for various leaching temperatures.Each slurry consisted of 5% solids.Figure E.18 shows the solution concentration of silicon for various initial concentrationsof sodium sulphide. It can be seen that as the initial concentration of Na2S increased, the Siconcentration in solution appeared to increase as well. This could be attributed to throughthe formation of NaOH in solution (through the dissociation of Na2S), allowing for more Sito be leached from the Pyrex beaker.284Appendix E. Additional Leaching FiguresFigure E.18: Silicon solution concentration after 24 hours for various initial concentrationsof sodium sulphide. Each slurry consisted of 5% solids.Figure E.19 shows the solution concentration of silicon for roasted and not roasted slags.It can be seen that the Si concentration appeared to marginally increase after roasting forslag A, while both slag B and the processed slag B reduced the Si solution concentrationafter roasting.285Appendix E. Additional Leaching FiguresFigure E.19: Silicon solution concentration after 24 hours for roasted and not roasted slags.Each slurry consisted of 5% solids.E.0.3 Aluminum Leaching BehaviourThe leaching behaviour of aluminum is examined in this section. BC Water Qualityguidelines for aluminum concentration were 0.1 mg/L for aquatic life, however all of the286Appendix E. Additional Leaching Figuresruns examined in this section had concentrations above this level [67]. As such, care mustbe taken to avoid spillage.Figure E.20 shows the aluminum recovery as a function of time for as-received slagsand the processed slag B. It can be seen that slag A had an extremely low Al recovery(0.62%), which correlated with an aluminum solution concentration of 1.3 mg/L, as seenin Figure E.21. Slag B and processed B saw Al recoveries between 15 and 20%, which ledto higher solution concentrations, as seen in Figure E.21. Although the processed slag Bhad a lower recovery of Al compared to the as-received slag B, the solution concentrationwas approximately double at 212 mg/L (compared to 111 mg/L for slag B). This increasewas due to leaching of alumina into the molten slag, as previously observed in the previouschapter, Section 5.2.3. The higher Al content of the processed slag B allowed for a lowerrecovery while having a higher Al solution concentration.Figure E.20: Aluminum recovery as a function of time for the as-received slags and theprocessed slag. Each slurry consisted of 5% solids.287Appendix E. Additional Leaching FiguresFigure E.21: Aluminum solution concentration after 24 hours for the as-received slags andthe processed slag. Each slurry consisted of 5% solids.Figure E.22 shows the aluminum recovery as a function of time for various slurry solidspercentages. It can be seen that similar to the previous elements examined, the recoveryof Al decreased as the solids content increased. Similar to the trend observed with Si, therecovery of Al for the 1% and 2.5% solids run was seen to reach a maximum value between 4and 8 hours, after which the recovery decreased. This was likely due to the precipitation ofaluminum species out of solution. Since both Al and Si had this trend, it could be impliedthat aluminosilicate species were likely formed in solution and precipitated.288Appendix E. Additional Leaching FiguresFigure E.22: Aluminum recovery as a function of time for various slurry solids percentages.Figure E.23 shows the concentration of aluminum in solution after 24 hours of leaching.It can be seen that as the solids concentration increased, the Al in solution increased. The1,062 mg/L Al observed in the 30% solids implies that the lower solids percentage runs didnot reach a solubility limit.289Appendix E. Additional Leaching FiguresFigure E.23: Aluminum solution concentration after 24 hours for various slurry solids per-centages.Figure E.24 shows the aluminum concentration in solution for 5 and 10% solids after24 and 96 hours of leaching. It can be seen that increasing the duration increased the Alsolution concentration by approximately four times for both 5 and 10% solids.290Appendix E. Additional Leaching FiguresFigure E.24: Aluminum solution concentration for 24 and 96 hour leach runs for 5 and 10%slurry solids.Figure E.25 shows the aluminum concentration in solution after 24 hours of leachingfor various initial NaOH concentrations. It can be seen that the addition of any amount ofNaOH resulted in an increase in the Al concentration in solution; however, values of 60 g/Land higher of NaOH had similar Al concentrations after the leach experiment concluded.This result implies that further additions of NaOH beyond 60 g/L did not have an effecton the leaching of Al after 24 hours.291Appendix E. Additional Leaching FiguresFigure E.25: Aluminum solution concentration after 24 hours for various initial concentra-tions of sodium hydroxide. Each slurry consisted of 5% solids.Figure E.26 shows the aluminum concentration in solution after 24 hours of leaching forvarious solution temperatures. It can be seen that a minor increase in the Al concentration,212 mg/L to 240 mg/L, occurred when the temperature was elevated from 25 to 50 ◦C.Further increases to the Al concentration did not appear to occur when the temperature292Appendix E. Additional Leaching Figureswas further elevated to 60 ◦C, as the concentration data points observed in the 60 ◦Crun overlapped those seen in the 50 ◦C run. At 50 ◦C, both the Si (Figure E.17) and Alconcentrations were seen to increase in solution.Figure E.26: Aluminum solution concentration after 24 hours for various leaching temper-atures. Each slurry consisted of 5% solids.Figure E.27 shows the concentration of aluminum in solution after 24 hours of leachingfor various initial concentrations of sodium sulphide. It can be seen that as the initialamount of Na2S increased, the Al concentration in solution appeared to increase in turn.293Appendix E. Additional Leaching FiguresFigure E.27: Aluminum solution concentration after 24 hours for various initial concentra-tions of sodium sulphide. Each slurry consisted of 5% solids.Figure E.28 shows the concentration of aluminum in solution after 24 hours of leachingfor roasted and not roasted slags. It can be seen that the roasting appears to have decreasedthe Al in solution by a minor amount for both slag B and the processed slag B.294Appendix E. Additional Leaching FiguresFigure E.28: Aluminum solution concentration after 24 hours for roasted and not roastedslags. Each slurry consisted of 5% solids.E.0.4 Arsenic Leaching BehaviourThe leaching behaviour of arsenic is examined in this section. Arsenic was an importantelement to consider as water quality guidelines dictate the maximum amount of arsenicpermissible in various water sources [68]. Similar to Se and Al, all of the experiments hadconcentrations above that for freshwater aquatic life (5 µg/L).Figure E.29 shows the arsenic recovery as a function of time for the as-received slagsand the processed slag B. It can be seen that both the as-received slags A and B had similarrecovery trends for As; however, the concentration of As was higher in slag B compared toslag A (Table 3.2), so the resulting higher recovery led to a higher concentration of As inthe leach solution, as seen in Figure E.30, ranging from 5.8 mg/L for slag A to 37.9 mg/Lfor slag B. The processed slag B saw a lower recovery than the other slags, and due to theprior pyrometallurgical processing step this slag had a lower As concentration in the slag,leading to a lower overall solution concentration.295Appendix E. Additional Leaching FiguresFigure E.29: Arsenic recovery as a function of time for the as-received slags and the pro-cessed slag. Each slurry consisted of 5% solids.296Appendix E. Additional Leaching FiguresFigure E.30: Arsenic solution concentration after 24 hours for the as-received slags and theprocessed slag. Each slurry consisted of 5% solids.Figure E.31 shows the arsenic recovery as a function of time for various slurry solid per-centages. Unlike the previous elements examined, the arsenic recovery to solution increasedas the amount of solids increased. While the recovery rates were below 25% for all runs,solution concentrations of up to 76 mg/L were recorded, as seen in Figure E.32.297Appendix E. Additional Leaching FiguresFigure E.31: Arsenic recovery as a function of time for various slurry solids percentages.298Appendix E. Additional Leaching FiguresFigure E.32: Arsenic solution concentration after 24 hours for various slurry solids percent-ages.Figure E.33 shows the concentration of arsenic for 5 and 10% solids percentages ofroasted and not roasted processed slag B. It can be seen that similar to other elementspreviously examined, the arsenic concentration increased when the leach was allowed torun from 24 hrs to 96 hrs.299Appendix E. Additional Leaching FiguresFigure E.33: Arsenic solution concentration for 24 and 96 hour leach runs for 5 and 10%slurry solids.Figure E.34 shows the concentration of arsenic in solution after 24 hours of leaching forthe various initial solution concentrations of sodium hydroxide. As the amount of the initialNaOH increased, it can be seen that the As concentration would subsequently increase, untilapproximately 60 g/L of NaOH were added. Values for As concentration at 60, 80, and100 g/L NaOH were seen to overlap, which could imply that there was no change in the Asconcentration between these values.300Appendix E. Additional Leaching FiguresFigure E.34: Arsenic solution concentration after 24 hours for various initial concentrationsof sodium hydroxide. Each slurry consisted of 5% solids.Figure E.35 shows the arsenic solution concentration after 24 hours of leaching for varioussolution temperatures. Compared to the baseline run, all the elevated temperature runssaw higher As concentration at the end of the run. It can be seen that the 50 ◦C run sawthe highest averaged As concentration, followed by 40 ◦C; however, the As concentration301Appendix E. Additional Leaching Figuresvalues observed in the 40 ◦C run had a larger deviation between them compared to theother runs. As such direct comparisons between 40 ◦C and other temperatures besidesthe baseline cannot be definitively stated. When comparing the 50 and 60 ◦C runs, itcan be seen that when the temperature was further increased that the As concentration insolution decreased, implying that a temperature range around 50 ◦C led to the highest Asconcentration.Figure E.35: Arsenic solution concentration after 24 hours for various leaching temperatures.Each slurry consisted of 5% solids.Figure E.36 shows the solution concentration of arsenic after 24 hours of leaching forvarious initial concentrations of sodium sulphide. Similar to previous elements, the As insolution was found to increase as more Na2S was added, however the increase was not aspronounced as some of the other elements observed, such as Se or Te. Interestingly, itappeared that the arsenic concentration decreased when sufficient quantities of Na2S wereadded; however, since only one data point exists for each of the Na2S runs, no exact claimscan be drawn from this observation. This decrease could be a statistical outlier, due to thetrend observed in the lower Na2S concentration runs not continuing, or potentially sufficientchanges occurred within the solution chemistry, either depressing the leaching of the arsenic,or potentially involving the precipitation of arsenic species from solution once a sufficientamount of Na2S had been added.302Appendix E. Additional Leaching FiguresFigure E.36: Arsenic solution concentration after 24 hours for various initial concentrationsof sodium sulphide. Each slurry consisted of 5% solids.Figure E.37 shows the arsenic concentration in solution after 24 hours of leaching forroasted and not roasted slags. It can be seen that the roast increased the As concentrationof each of the slags following the leach. This may be attributed to the oxidative propertiesof the roast converting arsenic species to a more soluble form.303E.1. Other Element BehaviourFigure E.37: Arsenic solution concentration after 24 hours for roasted and not roasted slags.Each slurry consisted of 5% solids.E.1 Other Element BehaviourIn addition to the elements examined in detail, the following five elements were also ofinterest in the leaching of the slag. However, unlike the previously examined elements, thefollowing five had simple or unremarkable leaching behaviour which could be easily summedup.SilverSilver was not observed in any of the leaching samples above values of 0.37 ppm. Assuch, it can be stated that the leaching process did not recover silver from the residue.CopperThe highest observed recovery for copper was 0.5%. In runs where the copper wasdetected, the variance between the data points did not allow for a trend to be observed.Values would fluctuate between time steps, again not allowing for a trend to be observed.The majority of runs had 0% Cu recovery after 24 hours. From these results, it was304E.1. Other Element Behaviourconcluded that the leaching process did not recover Cu into solution, and any reagents orchanges to the leaching parameters had no effect on Cu.BismuthSimilar to copper, it was concluded that the leaching process did not recover bismuth.The highest recovery values obtained were 2.5%, which decreased at the subsequent timestep, and several tests saw 0% Bi recovery at every time interval.LeadThe highest lead recovery obtained was 0.4%. As such, similar to the previous elements,the leaching process was not suitable for the recovery of lead, and the use of reagents andchanges to the leaching parameters had no identifiable impact on the Pb leaching. Wherelead was detected, no discernible trend could be determined due to fluctuations in the dataseries and variance between the data points at the same time step.AntimonyLastly, antimony, unlike the previous elements, was leached during the runs with suffi-cient recovery to be relevant in the subsequent electrowinning step. Data for Sb recoverywas obtained through mass balances using the slag residue. However, no discernible trendcould be identified for the recovery of Sb. The final recovery values would fluctuate between40 and 50 %, leading to concentrations of 19 to 22 mg/L Sb, with no trend observed be-tween runs. Therefore, while the leaching of Sb did occur, no reagent or change in leachingparameters appeared to have an effect on Sb leaching.305Appendix FFull Quadratic ModelsF.1 Thermodynamic Model FittingThe following tables are the MATLAB® output from the fitlm function for a fullquadratic model. X-values are X1 - temperature, X2 - PbO addition, X3 - Na2CO3, X4 -roasting temperature.Table F.1: Full quadratic model output from MATLAB® for the thermodynamic CCDmodel showing the coefficients for silver recovery to the metal phase.Estimate(Intercept) 99.7048X1 0.62342X2 0.15417X3 1.0329X4 0.011102X1X2 -0.085213X1X3 -0.25877X1X4 0.0098554X2X3 -0.1198X2X4 -0.015172X3X4 0.01044X12 -0.4343X22 0.070575X32 -0.64639X42 0.08657R2 0.867R2adj 0.782306F.1. Thermodynamic Model FittingTable F.2: Full quadratic model output from MATLAB® for the thermodynamic CCDmodel showing the coefficients for tellurium recovery to the slag phase.Estimate(Intercept) 99.9476X1 0.03597X2 -0.050403X3 0.1962X4 0.0003709X1X2 0.0049555X1X3 -0.019363X1X4 0.00058549X2X3 0.029569X2X4 -0.00086999X3X4 0.00033328X12 -0.060935X22 0.020064X32 -0.13275X42 0.018743R2 0.815R2adj 0.697307F.2. Pyrometallurgical ModelsF.2 Pyrometallurgical ModelsTellurium Recovery to the Slag PhaseTable F.3: Full quadratic model output from MATLAB® for the pyrometallurgical testingmodel showing the coefficients for tellurium recovery to the slag phase.Estimate(Intercept) 0.88227X1 0.0073017X2 -0.0037043X3 0.0006649X4 -0.00026389X5 0.0059447X1X2 -0.003828X1X3 0.009705X1X4 -0.0016747X1X5 0.0018629X2X3 0.0057945X2X4 -0.0028908X2X5 -0.0087541X3X4 -0.010864X3X5 -0.012372X4X5 -0.002528X12 -0.004647X22 0.0019887X32 0.0059658X42 0.0059783X52 -0.0057774R2 0.331R2adj -0.505F vs Const. Model 0.396Model p-value 0.974308F.2. Pyrometallurgical ModelsTellurium Recovery to the Metal PhaseTable F.4: Full quadratic model output from MATLAB® for the pyrometallurgical testingmodel showing the coefficients for tellurium recovery to the metal phase.Estimate(Intercept) 0.0020937X1 -0.0053052X2 -0.0015043X3 0.0025399X4 0.0019942X5 0.0014631X1X2 0.0016736X1X3 -0.0027668X1X4 -0.0026541X1X5 -0.0019901X2X3 -0.0013544X2X4 -0.0017281X2X5 -0.0023023X3X4 0.0023083X3X5 0.0016805X4X5 0.0012766X12 0.0024879X22 0.0012663X32 -0.00014034X42 0.00028254X52 -0.0002033R2 0.937R2adj 0.857F vs Const. Model 11.8Model p-value 3.78e-06309F.2. Pyrometallurgical ModelsSilver Recovery to the Metal PhaseTable F.5: Full quadratic model output from MATLAB® for the pyrometallurgical testingmodel showing the coefficients for silver recovery to the metal phase.Estimate(Intercept) 0.79657x1 0.061628x2 0.0005219x3 0.087118x4 0.043104x5 0.017948X1X2 -0.01003X1X3 -0.0030414X1X4 0.015176X1X5 0.027638X2X3 0.0071927X2X4 -0.0036275X2X5 0.038698X3X4 0.011284X3X5 0.0023569X4X5 -0.0113X12 -0.035437X22 -0.0029304X32 -0.04307X42 -0.0080953X52 0.0034715R2 0.74R2adj 0.416F vs Const. Model 2.28Model p-value 0.0494310F.2. Pyrometallurgical ModelsTotal Impurity Grade of the Metal PhaseTable F.6: Full quadratic model output from MATLAB® for the pyrometallurgical testingmodel showing the coefficients for the total impurity grade of the metal phase.Estimate(Intercept) 0.013441x1 -0.0062115x2 -0.0015606x3 -0.0014474x4 0.0010453x5 0.0013637X1X2 0.0015273X1X3 -0.00070082X1X4 -0.002893X1X5 -0.0022049X2X3 -0.0014106X2X4 -0.00073997X2X5 -0.0018041X3X4 0.0025008X3X5 0.0010408X4X5 0.00097292X12 0.0030238X22 0.0011188X32 0.00084571X42 0.00072226X52 -0.00020525R2 0.939R2adj 0.862F vs Const. Model 12.2Model p-value 2.98e-06311
Thesis/Dissertation
eng
Mechanical Engineering
Vancouver : University of British Columbia Library
Attribution-NonCommercial-NoDerivatives 4.0 International
Graduate
Bichler, Lukas
Recovery and production of high-purity tellurium from silver soda slag
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