David Dreisinger


Relevant Thesis-Based Degree Programs


Great Supervisor Week Mentions

Each year graduate students are encouraged to give kudos to their supervisors through social media and our website as part of #GreatSupervisorWeek. Below are students who mentioned this supervisor since the initiative was started in 2017.


Prof. Dreisinger is always kind and supportive to students and the students' future career developments. For Prof. Dreisinger, the graduate students are not the employees to finish the tasks but more like the "life children". He is knowledgeable with abundant engineering experience. However, he is always nice and behaves equally to students without showing the absolute authority. When students have trouble, no matter if it is related to their studies or life, once he knows, he is always there to provide help and support as much as he can in the first time. He always supports, helps and guides the students to be good for the world and to keep innovating in scientific research. From him, it can be clearly seen that the perfect combination of teacher, researcher, supervisor, and wise, kind and humorous elder! It is great luck for me to be one of his students!

Fei Wang (2019)


Graduate Student Supervision

Doctoral Student Supervision

Dissertations completed in 2010 or later are listed below. Please note that there is a 6-12 month delay to add the latest dissertations.

A study of basic iron sulphate formation and control in the autoclaving of argentiferous sulphide gold ores (2023)

A pressure oxidation (POX)-hot cure (HC)-lime boil (LB) process is used to recover gold and silver from refractory sulphide ores containing minerals like pyrite and arsenopyrite. Pressure oxidation is used to oxidize minerals and liberate occluded precious metal particles. Iron that is present goes into solution and under acidic and high temperature conditions often precipitates as basic iron sulphate. Basic iron sulphate consumes excess lime during neutralization prior to cyanidation. Therefore, a hot cure stage is required to re-dissolve this precipitate. In the presence of silver, conditions in the hot cure favour the formation of silver jarosite, which is refractory to cyanidation and must be decomposed prior to cyanidation. This is done in the lime boil to produce cyanide-soluble silver hydroxide. This study was conducted to investigate parameters that affect the precipitates formed in each stage of the POX-HC-LB process. In the autoclave, at 225 °C, a range of ferric and acid concentrations were studied. It was determined that at low concentrations of free ferric and free acid, hematite was the favourable iron precipitate. Above a critical concentration of free acid, basic iron sulphate was favoured. Speciation modeling of solution complexes in the ferric-sulphate-water system pointed to the possibility of FeSO₄HSO₄⁰, FeHSO₄⁺² and/or FeSO₄⁺ as precursors to basic iron sulphate. Speciation models were also used to calculate ΔG° values of basic iron sulphate for the first time, at varying temperatures.Basic iron sulphate formed in the autoclave dissolved quickly in the hot cure and silver precipitation occurred slowly. Silver precipitation was limited in the absence of other cations, whereas in the presence of alkali salts, silver co-precipitated with alkali jarosites.Silver that precipitated by itself, was readily soluble in cyanide, whereas silver that co-precipitated with alkali jarosites, had low amenability to cyanidation. Alkali treatment of the jarosite phases using lime saw successful breakdown of alkali jarosites but limited silver recovery by cyanidation. It was hypothesized that conversion of jarosites to hydroxides through reaction with lime is diffusion-controlled by a calcium carbonate product layer. Treatment with sodium hydroxide as the alkalising agent saw higher silver recovery, and further supported this theory.

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Comparative solvent extraction studies of molybdenum and other impurity metals in sulfuric acid medium by cyanex® 600 and fentamine™ TA0810 (2023)

Molybdenum is widely used as an alloying element in the manufacture of steels and super alloys. The widespread use of the element encourages a growing need to develop extraction processes for recovery from both primary and secondary sources. The Cytec Solvay Group has developed a solvent extractant, CYANEX® 600 for the recovery of molybdenum from acidic solutions originating from both low grade (e.g., copper solvent extraction raffinates, smelter dusts and slags) and high grade (molybdenite) molybdenum sources. CYANEX® 600 is a purified form of CYANEX® 272 with the active component, bis(2,4,4- trimethylpentyl)phosphinic acid. Heap leaching of copper from ores followed by solvent extraction is practiced commercially and the raffinate solution after the copper solvent extraction has substantial levels of molybdenum along with minor impurity metal ions in acidic pH range. The extraction behavior of CYANEX® 600 towards molybdenum in sulfuric acid medium was evaluated. The uptake of molybdenum increased with the increase in concentration of CYANEX® 600. An organic solution of 0.1 M CYANEX® 600 extracted ~97 % of 0.04 M Mo (VI) at pH
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Iodine-assisted heap leaching of chalcopyrite: laboratory and modelling studies (2022)

Copper mine operations face increasing challenges of processing low-grade chalcopyrite ores with elevated concentrations of impurities. Heap leaching is considered to be an effective and economically viable technology for treating these low-grade, complex copper sulfide ores. The addition of iodide in ferric sulfate solution is found to significantly enhance the dissolution of chalcopyrite at ambient temperatures. The aim of this research was to better understand the key factors that control the reaction rate and the mechanisms by which iodine accelerates the leaching kinetics. This research aim was achieved by conducting experimental and modelling studies of the iodine-assisted leaching of chalcopyrite concentrate and ore. Specifically, a series of leaching tests of increasing scale, from 50-mL bottle to 1-L reactor to 1-m and 6-m column tests, were carried out under partially- or fully-controlled leaching conditions; these leaching tests data were used to develop a kinetic model that correlates the copper extraction with the leaching conditions; the solid surface properties were examined by XRD, MLA, and XPS to uncover the leaching mechanisms; the calibrated kinetic model was finally used to assess the effect of key design and operating parameters on the performance of chalcopyrite dissolution in heap leaching.The experimental results show that the redox potential is the primary factor controlling the chalcopyrite leaching by controlling the iodine speciation in solution. Either diiodine or triiodide is considered to be the active oxidant responsible for chalcopyrite dissolution depending on the solution potential. The solid surface examination shows that the elemental sulfur and iron precipitates formed during leaching did not hinder the dissolution of chalcopyrite and that pyrite was inert during leaching. The kinetic model developed could simulate the copper extraction as a function of solution potential, total iodide concentration, and temperature. The sensitivity tests with the calibrated kinetic model show that the performance of the iodine-assisted leaching process can be improved by increasing irrigation rate, ferric concentration, iodide concentration, and temperature. The experimental and modelling results obtained in the present study can guide the design and implementation of the iodine-assisted heap leaching of chalcopyrite at an industrial scale.

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A study of selenate removal from water by adsorption on an aged, carbonated Friedel's salt material with mass loss considerations (2020)

In this work, selenate removal from water was studied using a material from the layered double hydroxide family. This material is a mixture of two hydrocalumite phases, chloro-hydrocalumite [Ca₄Al₂(OH)₁₂]·[Cl₂·4H₂O] and chloro-carboaluminate hydrocalumite [Ca₄Al₂(OH)₁₂]·[Cl(CO₃)₀.₅·10.8H₂O]. This material is called aged-FS in this thesis. Aged-FS was obtained by the co-precipitation method followed by an aging process. The aging condition was initial pH 12.0, liquid/solid ratio 20, 65 °C, for 7 days. During aging, a phase transformation from chloro-hydrocalumite in the hexagonal crystal system to chloro-carboaluminate hydrocalumite in the monoclinic crystal system was observed. The aluminum hydroxide impurity was removed and katoite (Ca3Al2(OH)12) was formed. The formation of well-shaped hexagon crystallites and agglomeration of particles were observed by morphology and particle size studies, respectively. The carbonate content increased during aging from 1.15 to 3.00 wt.%. By aging, ~50% reduction in FS dissolution was recorded. By some modifications in nitrogen purging of the synthesis process, the carbonate content was reduced from 1.15 to 1.01 wt.%. For selenate removal studies, the effect of initial pH, liquid/solid ratio, time, temperature, agitation, selenate concentration, sulfate and carbonate interference, and container type were studied. The highest loading percentage of selenate and the lowest dissolution of aged-FS was achieved at initial pH 12.0, liquid/solid ratio 200, 30 °C, 24 h, and 150 RPM. A selenate loading isotherm was fitted to the Langmuir model with maximum loading of 71.1 mg Se/g. The intercalated carbonate in aged-FS did not interfere with selenate removal while significant interference was observed from carbonate in solution. The sulfate anion did not interfere with selenate adsorption. The Al dissolution from aged-FS in an adsorption test was suppressed using calcium hydroxide solution. The Al concentration reached to 0.07 ppm; a value lower than the maximum instantaneous value set by a BC guideline. Calcination of aged-FS improved the selenate loading. The lowest level of selenium in solution achieved using the calcined aged-FS (1.84 ppb) met the 2016 US EPA maximum discharge limit of lotic water for Se (
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Mineral carbonation to sequester CO2 with concurrent metal sulfidization (2019)

Global warming is an urgent issue all over the world and mineral carbonation for CO₂ sequestration is one of the best methods to permanently store CO₂ gas. To make the mineral carbonation process profitable, it is suitable to combine with metal sulfidization for valuable metal recovery by utilizing the carbonation process since there are increasing demands of valuable metals around the world as well. This dissertation established the theoretical system for the potentially successful development of mineral carbonation for permanent CO₂ storage and utilization for metal sulfidization. The fundamental mechanisms and kinetics of mineral carbonation of olivine have been elucidated. The most important factors affecting the mineral carbonation process were the temperature, CO₂ partial pressure (PCO₂), specific surface area, aqueous ionic strength (I) and addition of sodium bicarbonate. The effects of these parameters on the mineral carbonation of olivine have been quantified. The mechanism can vary under different conditions and mainly depended on PCO₂ and the aqueous I. The increase of the aqueous I and PCO₂ can help prevent passivation of the mineral carbonation reaction by preventing formation of a silica-rich layer or a uniform carbonate layer respectively. Once the aqueous I and PCO₂ are high enough, the mineral carbonation of olivine is always controlled by chemical reaction of the dissolution of olivine. Under the chemical reaction control, a quantitative kinetic model has been developed, which can be used to predict the mineral carbonation efficiency and also the requirements of carbonation conditions. It is possible to utilize the mineral carbonation process for recovery of the released valuable metal from the dissolution of olivine by in-situ sulfidization. The key for the in-situ metal sulfidization is to continuously supply sulfide ions in the mineral carbonation system in order to selectively convert the released valuable bivalent metal ions to recoverable metal sulfides.

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Selenium removal from waste waters by chemical reduction with chromous ions (2019)

Selenium can be released to the environment from both natural and industrial activities, resulting in increased selenium concentrations in surface and ground water. Even small concentrations of selenium can be toxic for many forms of aquatic life. Therefore, selenium contamination in the receiving environment is a key issue for many industries. Selenium speciation in solution plays an important role in its removal. Selenite (SeO₃²⁻, Se(IV)) and selenate (SeO₄²⁻, Se(VI)) are the most important inorganic selenium species which are generally found in water and known to be toxic. Relative to the selenate, selenite can be quite easily removed from solutions using various treatment methods such as chemical reduction, precipitation and adsorption by ferrihydrite salts. However, these methods are not efficient for selenate removal. Typically, chemically based treatment processes for selenate removal require an initial reduction of selenate to the lower oxidation states (e.g., selenite, H₂Se). Chromous ions have been known as a powerful reducing agent in the reduction of many organic compounds, oxides, and sulphide minerals and in many proposed novel hydrometallurgical processes. Therefore, there is a high potential for chromous ions to reduce selenate effectively.In this study, the fundamental and practical aspects of the selenate reduction by chromous ions as a novel method to remove selenate from waste waters was investigated mainly in sulfate media. At first, the stoichiometry of selenate reduction by chromous ions was studied. Secondly, the kinetics of selenate reduction by chromous ions was studied over the wide range of acidity, chromous concentration, temperature and ionic strength. The reaction order with respect to the concentrations of selenate, chromous ions and hydrogen ions and the general rate law equation were determined. Furthermore, the effect of sulfate ions on the selenate reduction rate at different ionic strengths was studied. Thirdly, the reaction mechanism responsible for the reduction of selenate by chromous ions was suggested. Finally, the removal of hydrogen selenide generated from the reduction of selenate with chromous ions was studied using three reagents. Additionally, a hydrometallurgy flowsheet incorporating chromous generation, selenate reduction, hydrogen selenide removal, and chromic precipitation units was proposed.

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The modes of gold loss in the calcium thiosulfate leaching system (2018)

The processing of gold is becoming more complicated due to the increasingly complex nature of the remaining gold-bearing ore bodies. This worldwide phenomenon is the driving force for the development of alternative technologies for the leaching and recovery of gold from so-called double refractory ores. Barrick Gold recently commercialized a unique calcium thiosulfate leaching plant to treat these problematic ores after an autoclave pre-treatment with simultaneous recovery of the dissolved gold-thiosulfate complex onto anion exchange resins. Nonetheless, this new process can experience unexpected losses in gold from solution. It is hypothesized that the reagent’s degradation products known as polythionates and various mineral additions could adversely affect soluble gold stability along with their known detrimental effect on gold recovery. This dissertation aims to understand the possible causes of gold losses by means of a thorough investigation of the effects of polythionate and mineral additions into synthetic calcium thiosulfate leaching solutions. A series of batch leaching experiments were subsequently conducted on actual preg-robbing ores to further elucidate the effects of polythionates on the stability soluble gold complexes in the calcium thiosulfate system. A subsequent study on the rates of polythionate species loading on the resin and their competitive loading behaviour was made and extended to include their effects on the displacement of gold from the resin. This was demonstrated in the form of gold loading isotherms tailored to the calcium thiosulfate leaching system. Ultimately, the processes of gold leaching from refractory ores and gold recovery by resin loading in the presence of polythionates were tied together in a last study to quantify their harmful outcomes on overall gold extraction.

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Kinetics of sulphide conversion to thiosulphate and gold dissolution study by electrochemical quartz crystal microbalance (2016)

The kinetics of the oxidation of pyrite sulphide to a thiosulphate reaction were studied using a pyrite concentrate oxidized in an alkaline medium at an oxygen overpressure between 10 and 40 psig at a temperature of 80ºC. Sodium hydroxide was used to neutralize the acid produced as a result of the pyrite oxidation. The focus of this study was the calculation of thiosulphate yield as a function of sulphide sulphur concentration of the pyrite in the feed. A single rate expression combining kinetic constants of all the metastable oxyanions was derived to predict the thiosulphate yield as a function of the known pyrite sulphide sulphur concentration under the experimental conditions that were adopted. [formula omitted] It was found that at 20 psig oxygen overpressure and a temperature of 80ºC, the initial rate of sulphide oxidation and thiosulphate yield was close to 0.08 mol/h and 0.015 mol/h, respectively, at pH values greater than 12. However, a shift from linearity occurred when the pH decreased below 12. It was observed in the experiments that any decrease in pH was accompanied by an increase in solution potential, which enhanced the rate of gold leaching provided sufficient thiosulphate remained in the solution. The oxidation and reduction of gold in thiosulphate as a function of applied potential was studied using cyclic voltammetry. The responses of cyclic voltammetry of gold electrodes in thiosulphate were compared under different conditions involving catalysts such as thallium and thiourea. It was found that the gold anodic current increases in the presence of catalysts which is more likely to be related to the adsorption/desorption phenomenon and not to the depletion of the thiosulphate ion at the reaction interface. Higher anodic current increased the surface gold concentration available for reduction. The large anodic-to-cathodic peak separation that ranged from 0.25V to 0.37V allowed most of the leached gold to diffuse away from the electrode. This indicates the electrochemically-irreversible character of this system even in the presence of catalysts. The relationship between the reduction peak current and concentration of gold thiosulphate in the bulk electrolyte was estimated based on a relevant equation for electrochemically-irreversible reaction.

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Uranous sulfate precipitation as a novel route to uranium purification in extractive metallurgy (2015)

Uranous sulfate can be crystallized from uranium(IV)-containing solutions by raising the temperature and adding sulfuric acid. Several important aspects of the process have never been investigated, however, making its successful application as a real-world extractive metallurgy technology far from certain. This dissertation addresses several fundamental questions surrounding the crystallization of uranous sulfate from acidic process solutions. The effects of various parameters on the solubility of uranous sulfate and the kinetics of its precipitation are demonstrated, including temperature, acid concentration, and agitation, based on the results from a series of bench-scale experiments. The effects of various impurities on the selectivity and efficiency of the crystallization process are also determined. Two new uranous sulfate x-hydrate polymorphs, the hexahydrate and the octahydrate, are characterized using single-crystal x-ray diffraction, vibrational spectroscopy, and chemical assay data, and an understanding of the conditions under which they form is developed. The thermal stability and decomposition characteristics of uranous sulfate tetrahydrate, hexahydrate, and octahydrate are demonstrated through fundamental thermodynamic calculations and through the examination of thermal analysis data. The fundamental kinetics of uranium(IV) oxidation in acidic solutions are quantified through the interpretation of experimental data under various conditions of acidity, temperature, and oxygen partial pressure. Finally, a hydrometallurgy flow sheet incorporating uranous sulfate precipitation is presented, and the viability of the complete process is demonstrated experimentally, including electrolytic reduction, precipitation, filtration, drying, and calcining. This work demonstrates that uranous sulfate precipitation is viable as a hydrometallurgical process technology, and that further work is justified.

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Kinetics study of selenium and tellurium removal with cuprous ion from copper sulfate-sulfuric acid solution (2014)

Dissemination of selenium and tellurium in pyritic ores and many of the sulphide minerals results in the contamination of pregnant leach solutions and electrolytes in hydrometallurgical treatment of sulphide ores and residues. In an effort to reduce the detrimental effects of selenium and tellurium, it has been of great interest to remove selenium and tellurium from contaminated solutions to lower levels than allowed in regulations, product specifications or process requirements. The selenium and tellurium content of the solution may be reduced into insoluble precipitates of copper selenides and tellurides using cuprous ion. Cuprous ion has uses as a reducing agent in hydrometallurgical applications and can be specifically used to remove selenium and tellurium ions from copper sulphate-sulfuric acid solutions. In this study, the chemistry and kinetics of the removal of selenium and tellurium from copper sulfate-sulfuric acid solutions by cuprous ion reduction and precipitation was pursued. At first, a study of equilibrium cuprous concentrations for the cupric–copper metal reaction was performed and an empirical function capable of predicting the saturated [Cu+] was suggested at different temperatures, cupric and sulfuric acid concentrations. Secondly, the kinetics of the selenium removal with cuprous was studied and the mechanism of the reduction reaction and the rate law was determined over a wide range of acidity and temperature. The effects of temperature, acidity, cupric concentration and ionic strength on selenium removal rate were also studied. Subsequently, rate constants as functions of temperature, acidity and ionic strength were suggested.Thirdly, the tellurium reduction chemistry and reaction kinetics by means of cuprous in a wide range of conditions were investigated. The mechanism of the reaction, rate constants and activation energies were also determined. Similarly to selenium, effects of temperature and acidity on tellurium removal rate were also studied and rate constants as a function of temperature and acidity were suggested.Selenium and tellurium reduction reaction times can be estimated at different acidities and temperatures using the suggested rate laws and rate constant functions.

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Vanadium: Leaching & Solvent Extraction (2014)

The fundamental understanding of vanadium hydrometallurgy was developed in three phases: vanadium (V) leaching, vanadium (III) oxidative leaching, and solvent extraction of vanadium (V&IV).In the first section, V₂O₅ leaching was studied in three steps. First, vanadium leaching and solubility of VO₂⁺ at different pH’s and temperatures were investigated in sulfuric acid. Secondly, the kinetics of vanadium leaching in pH 5 and pH 8 solutions, and the reductive leaching of vanadium pentoxide using sodium sulfite were studied. It was shown that the kinetics of acid leaching is rapid but suffers from low solubility of VO₂⁺ in solution. Thirdly, the shrinking sphere model was employed to analyze the kinetics of reductive leaching. In the second step, V₂O₃ oxidative leaching was studied from 30°C to 90°C in sulfuric acid. This study has also been done in three different sections. First, the kinetics of oxidative leaching using oxygen was investigated. It was shown that this oxidative leaching is chemical reaction rate controlled with an activation energy of 69 kJ/mol. In the next step, it was shown that the presence of ferric enhanced kinetics significantly. Finally, oxidative leaching using a constant ferric-ferrous ratio from 1 to 300 was studied. The addition of KMnO₄ solution to the leach reactor was found to be a suitable oxidant for controlling solution potential. The oxidation rate using the constant ferric-ferrous ratio was very sensitive to temperature, with a large activation energy (38 kJ/mol) and the rate was proportional to the Fe(III)/Fe(II) concentration to the power of 0.47. In the third part, purification of synthetic vanadium-containing solutions using the solvent extraction technique was investigated. Various solvent extractants have been tested for vanadium recovery from acid leachates. One of the biggest problems for purification of the vanadium solution is iron separation. Therefore, this research assesses selectivity of vanadium over iron. The extraction of vanadium (V&IV), iron (III&II) with phosphinic acid (CYANEX 272), phosphonic acid (IONQUEST 801), phosphoric acid (D2EHPA) and phosphine oxide (CYANEX 923) extractants is reported. In addition, the extraction reactions for vanadium (V) and (IV) extraction using CYANEX 923 and D2EHPA were also studied.

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An Ion Exchange Loading Correlation for the Simulation of Base metal Resin-in-Pulp Circuits (2012)

This thesis investigates the loading rate of nickel onto an iminodiacetic ion exchange resin, with the goal of gaining insight into operation of a base metals resin-in-pulp (RIP) circuit. A modified pH-stat method was used to generate loading data under infinite solution volume conditions. A comparison of several commercially available iminodiacetic resins was performed, using this method.The standard linear approach to equilibrium was found to fit well to the portion of loading under film diffusion control, but none of the standard engineering approximations could adequately describe the data from the intraparticle diffusion / exchange rate limited regime. A hybrid correlation was developed and was found to adequately describe single element loading of nickel, copper, and cobalt from synthetic solution. In order to know which of the two models to use at a given point in time, a modified Helfferich number was derived. This dimensionless number can be used to track when the resin bead switches from film diffusion control to intraparticle diffusion control. The experimental fit parameters of the film diffusion model and the hybrid correlation were used to successfully predict the results of batch experiments with varying solution concentration. The effectiveness of these equations were also assessed and verified through the operation of a five stage RIP miniplant, using synthetic solutions. Using these circuit models, various RIP circuit operating configurations were simulated. Results from these simulations suggest that all operating configurations and strategies have both advantages and disadvantages. A table summarizing the process sensitivity of a cascade circuit to various operating conditions has been generated.

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Enhancing the kinetics of pyrite catalyzed leaching of chalcopyrite (2012)

An investigation has been conducted into the kinetics of chalcopyrite leaching in acidic ferric/ferrous sulfate media at atmospheric pressure and low temperature in the presence of pyrite. It has been found that pyrite samples from different sources affect the rate of chalcopyrite leaching differently. Some pyrite samples accelerate the rate significantly while others have little or no influence. The effectiveness of pyrite has a strong correlation with the level of silver occurring in the pyrite.In this study, the use of silver-enhanced pyrite in the Galvanox™ process to improve the extraction of copper from chalcopyrite was investigated. The catalytic properties of pyrite have been improved such that all pyrite samples accelerate the rate of copper extraction regardless of their sources. Under appropriate conditions, complete copper extraction can be obtained in the presence of silver-enhanced pyrite in 10 to 15 hours. Silver-enhanced pyrite can also be effectively recycled with minimal loss of effectiveness.A comprehensive understanding of the mechanisms involved in the process of chalcopyrite leaching in the presence of silver-enhanced pyrite was developed. It has been found that the acceleration of the rate of chalcopyrite leaching is due to the galvanic interaction between pyrite and chalcopyrite particles. An important factor in any galvanic process is maintaining electrical contact between two minerals to ensure the transport of electrons from the anode to the cathode. As chalcopyrite leaching proceeds, layers of extremely high electrical resistivity form around chalcopyrite particles, which limit the transport of electrons from chalcopyrite to pyrite and inhibit the galvanic interaction between these two minerals. However, in this study, it has been shown that, in the presence of silver-enhanced pyrite, around 10% of the added silver gradually leaves pyrite during leaching and reacts with the elemental sulfur layer on chalcopyrite to form silver sulfide. Although the amount of silver sulfide is very low and the sulfur layer does not become very conductive, it does become conductive enough to allow the transport of electrons at a rate sufficient to support the leaching reactions at rates observed in the process. With a conductive path between pyrite and chalcopyrite, the galvanic interaction between these two minerals can now occur.

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Solvent extraction of copper and cyanide from waste cyanide solution (2010)

The potential use of two commercial extractants, LIX 7950, a guanidine derivative, and LIX 7820, a solvent mixture of quaternary amine and nonylphenol, for recovery of copper and cyanide from waste cyanide solution has been investigated. Low equilibrium pH favors copper extraction while a high molar ratio of cyanide to copper depresses the copper loading. It is confirmed that Cu(CN)²⁻₃ is preferentially extracted over Cu(CN) ³⁻₄ and CN- by the extractants.Solvent extraction of the mixture of metal cyano complexes shows a selectivity order as follows: Zn > Ni > Cu > Fe. The presence of SO²⁻₄ or S₂O²⁻₃ shows an insignificant effect on copper extraction while SCN- ions may potentially compete for the available extractant with copper cyanide species and thus depress copper extraction significantly. Both extractants exhibit an affinity sequence as SCN- > CNO- > CN-> S₂O²⁻₃. The selectivity order of different anions with the extractants can be explained by the interrelated factors including anion hydration, charge density, compatibility of the formed complex with the organic phase and the geometry effect.The extraction of Cu(CN)²⁻₃ with LIX 7950 is exothermic with an enthalpy change (ΔH°) of -191 kJ/mol. The copper extraction with LIX 7820 has little change when the temperature is varied from 25 °C to 45 °C. For both extractants, the loaded copper and cyanide can be stripped efficiently by a moderately strong NaOH solution. Further increase in NaOH concentration results in the formation of a third phase. The presence of NaCN can facilitate stripping of the loaded copper and cyanide by favoring the formation of Cu(CN)³⁻₄ in the stripping solution.The important findings suggest a possible solution to the separation of metal cyanide species and free cyanide in the cyanide effluent. Both extractants can be used in a SX circuit for pre-concentrating copper into a small volume of strip solution which can be further treated byelectrowinning, AVR, SART or similar processes to recover copper products and cyanide. The free cyanide will remain in the raffinate solution from solvent extraction circuit which allows for the potential recycling of the barren solution to the gold cyanidation process.

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Sulfur dispersing agents for nickel sulfide leaching above the melting point of sulfur (2009)

The effects of sulfur dispersing agents (SDAs) in the oxygen pressure leaching of nickel concentrate at medium temperature were investigated. Liquid sulfur-aqueous solution interfacial tensions and liquid sulfur-sulfide mineral contact angles were measured at 140ºC, 690 kPa overpressure by nitrogen, and 1.0 mol/L NiSO₄. The effects of SDAs including lignosulfonate, Quebracho, o-phenylenediamine (OPD), and humic acid were evaluated by the calculation of the work of adhesion in the liquid sulfur-sulfide mineral-aqueous solution systems. It was found that the sulfide mineral surface is sulfophobic at pH from 4.1 to 4.5 due to the hydrolysis of nickel (II) ions to nickel hydroxide and the deposition of nickel hydroxide on the mineral surface. These findings apply to four different sulfide mineral systems, including pentlandite, nickeliferous pyrrhotite, pyrrhotite, and chalcopyrite. Lignosulfonate, Quebracho, and humic acid were found to significantly reduce the work of adhesion indicating they should be effective SDAs. OPD is ineffective in changing the work of adhesion of sulfur on the mineral sulfides indicating that it is not a good candidate for sulfur dispersion. The adsorption behavior of SDAs, including lignosulfonate, Quebracho, OPD, and humic acid on elemental sulfur and on nickel sulfide concentrate was investigated. Lignosulfonate, Quebracho, and humic acid were characterized by their infrared spectra. The charge changes on elemental sulfur surface were characterized by the measurement of the electrokinetic sonic amplitude (ESA) in the absence or presence of SDAs. The adsorption of lignosulfonate on molten sulfur surface was calculated by the Gibbs Equation. The adsorption of lignosulfonate, Quebracho, and humic acid on the nickel concentrate was investigated at ambient temperature. The adsorption of lignosulfonate, Quebracho, and humic acid on the nickel concentrate was found to be monolayer adsorption, which was fitted to the Langmuir adsorption isotherm. Electrostatic interaction and ion-binding are the possible mechanisms for the adsorption of lignosulfonate and humic acid on the nickel concentrate. Quebracho is adsorbed on the nickel concentrate through hydroxyl and sulfonate groups. OPD cannot adsorb on the molten sulfur surface. OPD undergoes chemical change in aqueous solution in the presence of ferric at ambient temperature.Oxygen pressure leaching experiments were performed at 140 or 150ºC under 690 kPa oxygen overpressure. The particle size of the nickel concentrate was found to be an important factor in leaching. During the leaching of nickel concentrate with P₈₀ of 48 µm, the SDAs were believed to be fully degraded before nickel was fully extracted. At most 66% nickel was extracted in the presence of 20 kg/t OPD. Fine grinding (P₈₀ of 10 µm) was sufficient for 99% nickel recovery at low pulp density while at high pulp density, the nickel extraction increased from 95% to 99% with addition of SDAs. Based on the leaching results on a nickel concentrate sample (-44 µm), OPD had the effect of increasing the nickel extraction to about 99%, followed by Quebracho (83%), lignosulfonate (72%), and humic acid (61%). It is suggested that the oxidation product of OPD is effective in solving the sulfur wetting problem in leaching. 97% nickel was recovered in the presence of 5 g/L chloride ion. Chloride ion has an effect to enhance the performance of lignosulfonate under leaching conditions.

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Master's Student Supervision

Theses completed in 2010 or later are listed below. Please note that there is a 6-12 month delay to add the latest theses.

Kinetic investigation of rare earth leaching on Labrador Canada rare earth ore in acidic sulphate medium (2023)

The kinetics of sulphuric acid leaching of rare earth elements in a sample from the Foxtrot deposit in Labrador, Canada was investigated. The selected parameters were, particle size fraction of +38 –53 µm, +53 -75 µm, +75 -106 µm, and +106 -150 µm, sulphuric acid concentration of 0.5 M, 1.5 M and 3 M, temperature of 30, 60, and 90 ℃. The highest extraction efficiency for Ce, Nd and Y was 96.72%, 84.91% and 72.83%, respectively at the highest temperature, highest sulphuric acid concentration and smallest particle size fraction. The highest selectivity was achieved at +106 -150 μm, temperature of 90 ℃, and sulphuric acid concentration of 0.5 M. It was determined that the kinetic experimental data followed the Jander product layer diffusion kinetic model, a type of shrinking core model, for all three elements. The activation energies for Ce, Nd, and Y were found to be 56.19 kJ/mol, 57.09 kJ/mol, and 38.57 kJ/mol, respectively. After identifying the most suitable kinetic model and calculating the activation energies, the rate orders were calculated for the mathematical model. These values were used to develop semi-empirical mathematical models for each element. The results showed that Y was the least affected by acid concentration and particle size, while Ce and Nd were the most affected. The acid leaching process resulted in crack and porosity formation for certain minerals but no notable change in particle size which is in a good agreement with the Jander kinetic model. The high activation energies with product layer diffusion control suggested that a mixed kinetic model might be in control of the leaching mechanism.

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Solvent extraction of nickel from complexation leachate solution (2023)

Hydrometallurgy uses aqueous solution chemistry for the recovery of metals from various materials. Complexation leaching uses chelating agents to bind targeted metals. Solvent extraction may then be used to recover and separate the target metal. This work focuses on the solvent extraction of nickel from complexation leaching solution. LIX 84-I was selected as an extractant to effectively recover nickel from near neutral leach solution in the presence of the complexing ligand. The recovery of nickel is determined by the thermodynamic relationship between the metal-ligand and metal-extractant formation. The factors that influence the thermodynamic equilibrium were investigated. The phase ratio (A/O) was studied. Solvent extraction involves the mixing and contact of an aqueous and organic solution. The kinetics of solvent extraction are therefore very important to minimize the equipment size and time for phase transfer. Thus, reaction time and temperature were studied. The stability of the ligand in the solvent extraction process was also investigated.Results show that nickel ion/citrate ligand is suitable for the solvent extraction of nickel due to the ease of loading on the extractant LIX 84-I and the ability to fully strip nickel at room temperature. However, a portion of citrate ligand is lost during the contact. The recycling and reusing of citrate ligand in the leaching process are therefore not ideal. Similarly, the loss of nitrilotriacetic acid ligand during the solvent extraction at high temperatures is measurable and therefore of concern. In comparison, ethylenediaminetetraacetic acid, ethylenediamine-N,N‘-disuccinic acid, and tetrasodium glutamate diacetate ligands are more stable during the solvent extraction process, but the kinetics of nickel extraction are too slow to permit industrial processing. It is recommended that additional factors such as higher temperature are explored to improve kinetics of nickel extraction (within the limits of safety).

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The alkaline pressure oxidation of gold-containing pyrite as a pretreatment prior to gold leaching : fundamentals and new approaches (2023)

Due to the depletion of gold ore easily processed, the demand for processing high carbonate refractory gold ore is increasing worldwide, and accordingly, the interest in alkaline pressure oxidation (POX) is also increasing. This study was conducted with the aim of investigating the fundamental kinetics of alkaline pressure oxidation and testing new concepts to improve the pre-existing alkaline POX process. Magnesium carbonate (MgCO₃) was proposed as an alternative alkaline reagent instead of sodium carbonate (Na₂CO₃). Formation of kieserite (MgSO₄·H₂O) at high temperatures has been confirmed. Addition of MgCO₃ may have a positive effect on cyanidation by replacing a part of the iron oxide - calcium sulfate based product layer with kieserite, which is expected to re-dissolve at low temperature and leave pores in the oxidized solid. However, the direct addition of MgCO₃ showed limitations in terms of extent of sulfide oxidation and gold recovery. A further attempt was made to affect the POX chemistry by addition of magnesium sulfate as a pre-treatment. This method showed potential regarding reduction of calcium carbonate (CaCO₃) reactivity as well as formation of desired magnesium hydroxide precipitate in the MgSO₄ pre-treatment stage. Two stage tests (MgSO₄ pre-treatment followed by alkaline POX) verified that the magnesium sulfate addition to POX had a beneficial effect on the gold recovery (8-10% increased) with a concurrent increase of sulfide oxidation. The kinetic study of pyrite oxidation showed the reaction rate can be modelled with the shrinking core model (SCM), and pyrite oxidation is under chemical reaction control in the initial stage regardless of key parameters (particle size, temperature, and Po2). However, rate control mechanism shifts to product layer diffusion with the effect of build-up of product layer. The activation energy for pyrite oxidation in the starting period was calculated at 36.1 kJ/mol which implied the reaction is controlled by chemical reaction, whereas the activation energy with a longer time span dramatically decreases to 15.9 kJ/mol (thought to be due to product layer diffusion rate control). In addition, it was found that excessive Na₂CO₃ addition to the POX system could have a detrimental effect on pyrite oxidation.

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The fundamental study of impurity behavior in wet chlorination of decopperized anode slime (2022)

The copper industry has been facing challenges associated with continuous declining ore grade and elevated levels of impurities in copper ores. Such challenges have already significantly impacted copper smelting and refining. The situation is compounded by increasingly complex feed materials processed by copper smelters. A major concern is the accumulation of impurity elements in copper anodes and the anode slimes generated in electrorefining. This, in turn, affects the downstream treatment of anode slimes for recovery of precious metals and platinum-group metals (PGMs) as valuable byproducts. Wet chlorination is hydrometallurgical process that involves the use of H₂O₂ and HCl for recovering these metals. The present study investigated the leaching behavior of not only gold and PGMs but also a range of impurity elements in the wet chlorination using both thermodynamic calculations and laboratory wet chlorination tests.The stable species of all elements present in the decopperized anode slime were identified in the Pourbaix diagram (Eh-pH diagrams) constructed. The results generally agreed with those reported in the literature, even though some species reported were not identifiable due to a lack of relevant thermodynamic data. The major findings of the laboratory wet chlorination tests were summarized as follows. Antimony played an important role in the behavior of impurity elements. The oxidation of antimony from Sb(III) to Sb(V) occurred in the ORP range of 600 - 700 mV, causing the precipitation of Sb(V) together with arsenic and bismuth as arsenato antimonates. The extractions of arsenic, antimony, bismuth, and tin were initially independent of but later affected by the rate of H₂O₂ addition due to the removal of antimony (V) from solution via precipitation of arsenato antimonates. The extraction of arsenic, antimony, bismuth, and tin increased with the leaching temperature. The extractions of arsenic, bismuth, and tin were affected by acidity but not by the total chloride concentration. However, the antimony extraction was affected by not only the proton concentration but also the chloride concentration. Simple alkaline leaching without H₂O₂ addition was shown to be possibly effective to remove some impurity elements, especially tellurium, arsenic, and tin, prior to the extraction of gold and PGMs.

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Fluoride removal from brine with ion exchange resin (2020)

The presence of fluoride has caused a number of complications in various industrial applications. In the sodium sulfate salt splitting technology developed by NORAM Engineering and Constructors, the undesirable fluoride ions in brine solution could cause the breakdown of the protective passivation layer on the titanium anode, resulting in pit corrosion or spalling of the coating on the dimensionally stable anode (DSA). Conventionally, adsorbents such as activated alumina (AA), coagulation including calcium fluoride precipitation method, ion-exchange (IX), solvent extraction (SX), and reverse osmosis (RO) have been used to remove fluoride from solution. Of all the options, CaF2 precipitation method is the most commonly used to remove fluoride from solution. However, along with its slow nucleation, CaF₂ has a theoretical solubility limit of 8 ppm F- at stoichiometric concentration of calcium in wastewater [1]. Therefore, a different approach for a robust fluoride removal system is required. In this work, aluminum or zirconium was pre-loaded onto LANXESS Lewatit Monoplus TP 260 amino phosphonic acid functional group cation chelating resin. These resins were then used to selectively load fluoride from Na₂SO₄ brine solution. Preliminary batch isotherm studies revealed that aluminum can be readily loaded onto this resin type; however, the zirconium loading capacity was poor due to the solubility limit of zirconium where low pH inhibits an effective loading process. Subsequently, fluoride batch isotherm studies revealed that the maximum fluoride loading capacity of Al pre-loaded and Zr pre-loaded resin were 1.30 mol/kg Al-resin and 0.70 mol/kg Zr-resin from 12 wt.% Na2SO4 brine, respectively. However, in the column loading trials, an increase in fluoride loading capacity was recorded with Zr pre-loaded resin as the cycles of loading and regeneration continued. Most notably, Zr pre-loaded resin provided minimal metal leakage (Zr) during the sorption process while reducing the fluoride concentration below 0.5 mg/L throughout all the loading cycles.

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Thiocyanate leaching of gold (2020)

This study focused on evaluating sodium thiocyanate as an alternative reagent to the conventional cyanidation process for leaching of gold ores. Goldcorp’s Coffee project in Yukon-Canada supplied three mineral samples namely, Supremo Oxide 68151, Supremo oxide 68151, Supremo Composite 72142 for this study.As a baseline for comparison with thiocyanate extraction results, cyanidation tests performed on all the three samples showed that the samples are amenable to the conventional cyanidation leaching, yielding gold extractions as high as 97% for Supremo composite 72142.A series of leaching tests were performed on the 72142 B sample with SCN⁻ solutions to determine the feasible regions of gold dissolution and to maximize gold dissolution. The leaching tests were conducted in the acidic regime (pH 1.5 -2) for these samples. Notable results with SCN⁻, ferric sulphate and potassium iodide variation showed gold extractions of 91 % in solutions containing 0.15 M thiocyanate; 92% with 0.15 M SCN⁻ and 0.10 M Fe(III); 94 % with 0.10 M SCN⁻ and 0.05 M KI; 94 % with 0.10 M SCN⁻, 0.05 M Fe(III) and 0.02 M KI and 95 % with 0.15 M SCN⁻ and 10 g/L H₂O₂.The kinetic leach data were well fitted by the CIP/CIL leach model developed by Nicol et al, giving estimates of the leach rate parameter of 72.1 hr⁻¹ and leach tails grade after infinite leach time of 0.17 g/t, confirming fast leaching of the 72142 B sample.The tests ended with gold adsorption from thiocyanate solution onto carbon. The average gold loading onto carbon was 2100 g/t and 48 g/t for carbon concentrations of 0.25 g/L and 20 g/L, respectively. Results obtained were excellent with greater than 98 % gold adsorbed in less than 0.5 hr when carbon concentration was above 5 g/L.The results show that the gold ore from the Goldcorp’s Coffee project in Yukon-Canada is amenable to extraction with acidified thiocyanate based lixiviant and subsequent adsorption of the gold-thiocyanate complex onto activated carbon, giving gold extraction results that are comparable to cyanide-based gold extraction. The thiocyanate system is, therefore, a competitive and alternative leaching reagent to conventional cyanidation.

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Cupric ammoniacal thiosulfate leaching of natural acanthite ore in a continuous circuit (2019)

Cyanidation has been the conventional way to leach silver ore for many years. However, environmental concerns regarding the use and disposal of cyanide have increased public concern. Now, regulations have emerged that prevent or limit cyanide usage in many regions. Finding an alternative lixiviant for silver leaching is necessary to leach silver ores in these locations. Among the proposed alternative leaching agents, cupric-ammoniacal thiosulfate system is a very attractive solution. Although the research on cupric-ammoniacal thiosulfate leaching system has been conducted for many years, most of the experiments were performed in batch fashion. It is difficult to find scale-up leaching data or experiments.This thesis study aims to engineer a lab-scale continuous leaching system that uses cupric ammoniacal thiosulfate as the leaching agent. The goal is to mimic the industrial process and collect useful data forfurther research and application. The experiments started with a series of batch leaching tests to obtain the optimized leaching conditions for treatment of an ore sample containing acanthite received from PanAmerican Silver. Next, the optimized leaching conditions were applied to the continuous leaching tests. After leaching, two different types of recovery methods were studied to separate the silver from the pregnant leach solution. The two recovery methods are ion exchange and cementation. The tests results show that the leaching ability of cupric-ammoniacal thiosulfate system is good. The continuous leachingtests were run for 3 days with a 24-hour retention time in leaching and achieved an 86% silver leaching efficiency. Furthermore, the leached silver was easily recovered 100% by cementation. The ion exchangetests were not successful in demonstrating silver recovery from the leachate.

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Lithium extraction from brine with ion exchange resin and ferric phosphate (2019)

Lithium is an essential metal for our society. Notably, increasing energy storage system will necessitate much more lithium in the future. This study focused on brine deposit while lithium exists in hard rocks as well. Conventionally, solar evaporation has been used to concentrate lithium from brine, but it takes more than one year. Thus, a more rapid process is desired for the accelerating demand. Here, two types of adsorbent, ion exchange (IX) resin and heterosite ferric phosphate (FP), were studied in order to extract lithium selectively from brine rapidly. First, more than thirty IX resins were tested in lithium chloride solution. Out of the thirty, sulfonate, iminodiacetate and aminomethylphosphonate resins succeeded in extracting lithium with the value of 16.3–32.9 mg-Li/g. However, no resins could adsorb lithium from a mixed brine solution which contains other interfering cations like sodium. An aluminum loaded resin was also tested since some past studies had reported lithium selectivity with this material. Its adsorption density was 6.6 mg-Li/g and was higher than any other resins tested for the mixed brine in this study. Nevertheless, the overall results showed that the IX resins were not so suitable for lithium extraction from a mixed brine. Then, heterosite FP was investigated as an alternative adsorbent. The FP can adsorb lithium selectively with the addition of a reducing agent to form lithium iron phosphate. This study used thiosulfate (TS) and sulfite (SF) individually as a reducing agent. The maximum adsorption density was 45.9 mg-Li/g by SF reduction at 65 °C, which is almost the same as the theoretical value of 46.0 mg-Li/g. The maximum selectivity over sodium was 2541 by SF reduction at 45 °C. Additionally, it was confirmed that the FP could be recycled by persulfate oxidation without degradation. Finally, the kinetics was studied and fit using pseudo first-order and shrinking sphere model. The two models fit the experimental results and indicated that the lithium extraction reaction was chemical reaction controlled. Since the FP method was found to be promising, it is highly recommended that it should be developed further by using natural brine sources.

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Atmospheric alkaline pre-oxidation of refractory sulphide gold ores (2018)

Four iron-sulphide-containing gold ores from the White Mountain mine in the Jilin Province, China were studied to examine the benefits of using oxidative treatment prior to conventional cyanidation to leach gold. This pre-treatment consisted of bubbling air into an ore/water mixture for eight hours while adding enough caustic sodium hydroxide to maintain a pH of 11 throughout the test.Three physical parameters were studied: total consumption of sodium hydroxide, rate of slurry aeration, and temperature of the slurry. Cyanide consumption was studied to determine whether increased oxidation of the ore would result in excessive degradation of sodium cyanide or otherwise impact the level of gold extraction. X-Ray Diffraction analysis was conducted on untreated samples as well as the final residues from the end of select tests to investigate by-products that precipitated after oxidation. Finally, an indicative economic analysis was conducted for each ore sample by comparing the increased dosage of caustic reagent applied during pre-treatment with the corresponding increased gold extraction.Results of this study show that aerated pre-oxidation of three of the four ores tested increased gold extraction compared to conventional cyanidation. Higher caustic dosages may have led to increased oxidation and were symptomatic of elevated release of sulphuric acid due to sulphide oxidation. Gold extraction was seen to rise in most cases after an increased rate of aeration during pre-treatment. Finally, the iron-sulphide ores tested responded extremely favourably to raised temperatures, showing both increased iron and sulphur oxidation, as well as higher gold extractions.Major losses of cyanide, likely due to production of cyanate and thiocyanate, were observed. Residual cyanide levels in the leach remained positive and did not affect gold extractions. Iron oxides such as hematite, and magnetite, and iron oxy-hydroxides such as goethite and ferrihydrite were observed in oxidized ore residues, indicating a high level of iron and sulphur oxidation.

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The use of surfactants to aid and improve the leaching of low grade copper ores (2018)

Copper heap leach operations often suffer from reduced efficiency due to long leach times and variable recoveries. Surfactants have been considered as an option in increasing the leachability of ores. Improvements in overall copper extraction have been noted with their use, though testing has only been conducted on a limited scale. The molecular function of surfactants in heap leaching has not been extensively studied and is not well explored. The work in this thesis was aimed at better understanding and characterizing the function of the surfactants. Work was performed with surfactants developed by BASF specifically for heap leaching.Initial experimentation consisted of using flooded vats to compare copper extraction from ores. Leach solution with and without surfactant was fed to the ores. The presence of surfactants was noted to increase the overall copper recovered by approximately 2-3%. Interfacial tension measurements were performed to determine the changes imparted onto the acidic leach solution by the surfactants. Hanging drops were used to determine the activity at the air-liquid boundary. It was found that at the surfactant concentrations used in heap leaching, the interfacial tension of the fluid changed very little, from about 71 mN/m to 69.5 mN/m. The contact angle was determined to better understand the interaction between the acidic media and the ore. This was obtained using capillary wicking and Washburn’s equation. Ore was finely ground and packed into particle beds. Leach liquid with surfactant was introduced to these beds. The rate of permeating fluid flow was monitored against time. The affinity of the liquid for the solid surface dictated the rate of uptake. Washburn’s equation allowed for the contact angle to be calculated from these results. It was found that surfactants lowered the contact angle of liquid on solid by up to 3 degrees. The combination of results indicated that the surfactants increases the affinity between the solid and liquid by reducing the contact angle. In a heap, this allows acid to ingress further into sub-surface regions of ore particles. As a result, leachability of the ore is increased as harder to reach minerals can be accessed.

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Thiosulfate leaching of natural Acanthite ore in copper-ammonia-ammonium sulfate medium (2017)

Silver is commonly present in acanthite in nature. Reagents like cyanide are used to extract silver from acanthite ores. However, cyanide can potentially damage human health and environment. The use of cyanide is tightly regulated, thus forcing the industry to seek for alternatives. Thiosulfate is currently the most promising alternative. The leaching chemistry of silver with thiosulfate is complex and maybe supplemented with additives such as ammonia, copper and even ethylenediaminetetraacetic acid. The efficiency of silver leaching is improved with the use of these additives. The use of cyanide for silver leaching in Navidad project in Argentina is not permitted, so the use of thiosulfate leaching as an alternative was investigated. The application of thiosulfate leaching to Navidad ores containing acanthite was the focus of this thesis. This thesis provides experimental evidence that supports the use of thiosulfate with additives as a promising alternative to conventional cyanidation method for the Navidad deposits and for similar deposits, wherever found.Thiosulfate leaching of silver is known for two pathways: silver in acanthite is substituted by cupric or by cuprous ion. The cupric pathway is thermodynamically more favourable, but various factors may affect extraction. Batch leaching tests showed that Navidad ore samples may be leached using thiosulfate, with silver extraction affected by variables including thiosulfate concentration, ammonia concentration, initial copper addition, pH, temperature, EDTA addition and the presence or absence of air sparging. The most significant variables were thiosulfate concentration, ammonia concentration, copper addition and pH. Cyanidation yielded 91.2% extraction of silver from a sample of Loma de la Plata, and thiosulfate leaching with 0.2 M of thiosulfate and 1.0 M of ammonia yielded comparable extractions: 92.1% and 87.0%, respectively. Initial copper addition increases extraction rate from 66.2% to 72.3% after 72 hours, and air sparging increases extraction rate to 84.8% after 72 hours. Other samples from the Navidad Project were also tested and found to be amenable to thiosulfate leaching. LDLPMC and Connector Zone (CZMC) sample were found to have potential for thiosulfate leaching to achieve a high silver extraction.

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Electrorefining of High Purity Manganese (2016)

Manganese is an important industrial metal used as an additive for production of various steels, non-ferrous alloys, electronic components and special chemicals. The traditional electrowinning process is not suitable for the production of high-purity manganese. Therefore, a novel ion exchange - electrorefining process for production of high-purity manganese is proposed to solve the problems. The ammonium chloride – manganese chloride electrolyte was selected as the most promising for refining. The physical properties of the electrolyte were first studied. The density increases when increasing concentrations of the ammonium chloride and manganese chloride. The maximum electrical conductivity is obtained with high concentrations of ammonium chloride and high temperature. Viscosity is minimized by high temperature and by low concentrations of manganese chloride. The electrorefining process introduces impurities into the anolyte from the dissolving anode. The purification of the electrolyte by cementation and ion exchange were investigated. This thesis reports the uptake of copper, nickel, cobalt, cadmium, zinc and manganese from manganese chloride solution onto the chelating resin Lewatit® MDS TP220 and Purolite S930Plus in batch and column experiments. The results demonstrate the ability for Lewatit® MDS TP220 to remove contaminants to an extent satisfying the quality criteria required for electrorefining. However, cementation of impurities by manganese powder and ion exchange with Purolite S930Plus are not suitable for purification to satisfy the quality criteria. In the electrorefining process, the individual and synergistic effects of selected impurities on manganese deposit quality were first investigated. The purity of manganese deposit did not change significantly in the presence of impurities. An addition of 0.15–5.0 mg/l Zn²⁺ to the catholyte increases the cathodic current efficiency. The parameters of current density, deposition time, and cathode usage frequency were investigated. Current density, deposition time and cathode usage frequency could affect current efficiency, specific energy consumption and surface morphology. With increasing the current density, the cathodic current efficiency first increased, reached a maximum value and then decreased. With increasing deposition time from 24 hours to 48 hours, the manganese current efficiency decreased and the deposit became more dendritic. The more times the cathode was used, the lower the current efficiency.

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Fundamental study on tin recovery in acidic aqueous systems (2015)

Tin is widely used in solder, tin plating and tin alloys. The current recovery rate of tin metal is low and insufficient with just over 300,000 tonnes annually. The grade of tin concentrates in traditional smelting methods needs to be at least 60%. Otherwise, iron, the chief impurity in tin concentrates can form tin-iron alloys and result in inefficient recovery of tin. Therefore, a hydrometallurgical technology to treat lower grade tin concentrates is proposed to solve the problem and close the demand gap for tin. The electrochemical reduction of chromium(III) solutions was conducted with a graphite felt cathode in acidic aqueous systems (chloride, sulfate and MSA). The parameters of acid concentration, current density, graphite felt thickness, graphite felt surface condition and graphite felt usage frequency were investigated. It was found that acid concentration has a significant influence on chromium(III) reduction in the sulfate and MSA system, while slight effect in the chloride system. In addition, the lifetime of the graphite felt in the sulfate and MSA system was shorter than that in the chloride system. These electrochemical differences may result from the pathway difference in electron transfer between the electrode and the chromium(III) ions. In general, chromium(III) ions in the chloride system showed the best electrochemical reduction activity. The chromium(II) ions synthesized from electrochemical reduction of chromium(III) ions were then used to effect the reduction of SnO₂ powder. The effect of temperature on the recovery test in the chloride, sulfate and MSA system was investigated. It was found that under the conditions of this thesis, the predominant recovery product of SnO₂ is Sn metal, rather than Sn(II). Generally, the recovery kinetics and total conversion were low in the sulfate and MSA system; however, the chloride system showed significantly better recovery results. This may be attributed to the catalysis effect of the chloride ions on the recovery process.

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Kinetics of Leaching of Covellite in Ferric-Sulfate-Sulfuric Acid Media (2015)

Hydrometallurgy methods to extract copper are becoming more frequently applied in modern industry. However, the leaching kinetics for certain minerals like covellite is poorly understood. This thesis investigates the kinetics of covellite leaching in a ferric-sulfate-sulfuric acid media, with an emphasis first placed on the understanding of the effect of the most common variables such as temperature and redox potential. A natural mineral sample of covellite and an ore sample from the Oyu Tolgoi project in Mongolia were obtained for the leaching studies. The leaching temperature was varied from 20° to 90°C, the total iron concentration varied from 0.1 mol/L to 0.5 mol/L, the Fe+³/Fe+² ratios varied from 0.1 to 10. The leaching results showed that an increase in temperature will result in an increase in the rate and extent of copper extraction. However, the redox potential or Fe+³/Fe+² ratio have little to no effect on the final copper extraction. These factors had only a modest impact on copper leach kinetics. The final copper extractions for covellite from Butte, Montana and covellite containing ore from Oyu Tolgoi at the same temperature were very similar. The key factor to improve copper extraction from covellite containing ores is to maximize the leach temperature. The other factors appear to be much less important. These findings provide the basis for process design and optimization of industrial leaching processes of covellite.

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Application of an ion exchange loading correlation for nickel recovery from a ferrous containing solution (2014)

Nickeliferous laterite ore is one of the resources for nickel recovery that can be processed by using commercially-available processes such as High Pressure Acid Leach. Since this is a costly hydrometallurgical process, maximum nickel recovery is required to justify the investment. Recently, the use of ion exchange resins and resin-in-pulp processes have been considered and proposed as methods of primary or secondary metal recovery from nickel laterite leach slurries and leach tailings. The solution produced in the high pressure acid leach treatment of laterite ores contains several impurities. As a result, the presence of high levels of impurities such as the ferrous ion is expected to be present in an actual resin-in-pulp feed slurry. Many of these impurities (including ferrous) initially load onto the resin and are then displaced by the target metal ion (nickel). Thus, it is expected to see a different nickel loading manner in the presence of ferrous ions compared to the loading of nickel in the absence of any impurities (i.e. nickel load onto resin only by displacing hydrogen ions). This emphasizes the significance of testing the desired ion exchange resin under conditions reflective of the expected industrial operating conditions.This thesis reports on the investigation of the loading rate of nickel onto an iminodiacetic acid ion exchange resin (TP207XL) under the finite solution volume condition in the presence of ferrous in a batch loading system. Initially, the batch loading rate of the ferrous ion onto resin under the infinite solution volume condition was examined. A hybrid correlation model was applied for both ferrous loading and nickel-ferrous displacement loading based on the kinetics experimental datasets. The correlation was found to adequately fit the ferrous loading datasets. According to the obtained fit parameters, it was proven that the loading rate of the ferrous ions onto the resin is slower than that of copper, nickel and cobalt ions. Furthermore, in the case of ferrous-nickel displacement tests, the experimental fit parameters of the hybrid correlation were used to successfully predict the results of the batch loading experiments by applying the finite difference model.

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Comparative study of auxiliary oxidants in the cyanidation of silver sulfide (2014)

Noticeably lower extraction of silver than gold in conventional cyanidation process has been commonly observed. Acanthite (Ag₂S) is one of the typical silver minerals with low solubility in aqueous medium, which result in low overall silver extraction. This thesis discusses the performance of hydrogen peroxide (H₂O₂), calcium peroxide (CaO₂), lead nitrate (Pb(NO₃)₂) and LeachWELL™ as potential auxiliary oxidants to improve the conventional cyanidation process. The dissolution of synthetic silver sulfide in cyanide solution with different oxidants was studied with stirred reactor leaching tests. The addition of H₂O₂ increased final silver extraction, and also raised cyanide and hydroxide consumption. The addition of CaO₂ had few effects on silver extraction. LeachWELL™ had a better performance than H₂O₂ and CaO₂ in terms of increasing silver extraction and reducing reagent cost. Furthermore, Pb(NO₃)₂ was also tested on synthetic silver sulfide leaching separately and was shown to have a similar effects as LeachWELL™ on this system. The effect of leaching parameters in the LeachWELL™-cyanide system, such as LeachWELL™ concentration, cyanide concentration, temperature and pH, were also studied. It was indicated that the increase of LeachWELL™ and cyanide concentration could further accelerate leaching and the performance of LeachWELL™ in this system was not affected by pulp pH. The extraction of silver from synthetic silver sulfide in the LeachWELL™–cyanide system was influenced to different extents by the presence of non-silver sulfide minerals.In general, most of the non-silver sulfides interfered with silver extraction via consumption of LeachWELL™ and cyanide. However, galena acted as a promoter and enhanced silver extraction.The performance of H₂O₂, CaO₂ and LeachWELL™ on natural acanthite sample cyanidation was also tested. Leaching of the natural lead-bearing acanthite sample with no auxiliary oxidant presented higher silver recovery than that achieved when leaching synthetic Ag₂S. The addition of H₂O₂ or CaO₂ further accelerated leaching while the addition of LeachWELL™ had no significant effect. It could be concluded that with enough lead in the leaching system the presence of sufficient oxidant could further enhance cyanidation of acanthite.

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Fundamental study on extracting lead from cerussite concentrate in methane sulfonic acid based solution (2013)

Cerussite (PbCO₃) concentrates may be recovered from oxidized lead ores. These concentrates suffer from the intrinsic fuel shortage via traditional smelting route and may lead to widespread contamination in shipment. Therefore, a hydro-electrometallurgical process to treat cerussite concentrate by methane sulfonic acid (MSA) is proposed to solve the above problems. The leaching of cerussite concentrate by MSA was first studied. The parameters of stirring speed, temperature, acid concentration, particle sizes and solid concentration were considered. It was found that stirring speed, temperature, particle size and proton concentration had a significant influence on the kinetics while solid concentration showed no effect on the final lead extraction under the experimental conditions. The leaching results indicated that MSA is favourable to treat cerussite concentrate and the lead extraction could achieve the theoretical maximum in just 10 min at ambient conditions. The lead content in solution obtained at high solid concentration was sufficiently high, and easily met the concentration requirements for the subsequent electrolysis process. After leaching, the residue was subjected to a desulfurization treatment to recover the remaining lead in residue. Using desulfurization agent Na₂CO₃, the remaining lead, mainly in anglesite, was transformed to PbCO₃ that was followed by a re-leaching treatment with MSA. An overall lead recovery of 98% was finally obtained. In the electrowinning process from MSA based electrolyte, the individual and synergistic effects of two ligninsulfonate salts and two glycol-type agents on lead deposit quality were first investigated. Compared to the other three additives, the individual use of calcium ligninsulfonate most benefited the morphology of lead deposit. The operating parameters in the lead electrowinning process (i.e. temperature, current density, concentrations of lead ion and protons) had a widely acceptable range. The cathodic current efficiency and specific energy consumption in most tests were around 99% and 0.53Wh/kg, respectively. The SEM micrographs showed that the lead deposits obtained under the optimal conditions were compact and even. These results for leaching and electrolysis made the MSA system competitive to the comparable fluoborate and fluosilicate systems. Finally a simplified flowsheet to extract lead from cerussite concentrate in MSA based solution was proposed.

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Mecury stabilization using thiosulfate or selenosulfate (2013)

Mercury is often found associated with gold and silver minerals in ore bodies. It is recovered as liquid elemental mercury in several stages including carbon adsorption, carbon elution, electrowinning and retorting. Thus a great amount of mercury is produced as a by-product in gold mines. The Mercury Export Ban Act of 2008 prohibits conveying, selling and distributing elemental mercury by federal agencies in United States. It also bans the export of elemental mercury starting January 1, 2013. As a result, a long-term mercury management plan is required by gold mining companies that generate liquid mercury as a by-product. This thesis will develop a process to effectively convert elemental mercury into much more stable mercury sulfide and mercury selenide for safe disposal. The process consists of 1) extraction of elemental mercury into solution to form aqueous mercury (II) and 2) mercury precipitation as mercury sulfide or mercury selenide. Elemental mercury can be effectively extracted by using hypochlorite solution in acidic environment to form aqueous mercury (II) chloride. The effect of different parameters on the extent and rate of mercury extraction were studied, such as pH, temperature, stirring speed and hypochlorite concentration. Results show that near complete extraction can be achieved within 8 hours by using excess sodium hypochlorite at pH 4 with a fast stirring speed of 1000RPM. Mercury precipitation was achieved by using thiosulfate and selenosulfate solution. In thiosulfate precipitation, cinnabar, metacinnabar or a mixture of both can be obtained depending on the experimental conditions. Elevated temperatures, acidic environment and high reagent concentrations favour the precipitation reaction. Complete mercury removal can be achieved within 4 hours. However, it appears that the less stable metacinnabar tends to form when the precipitation rate increases. Selenosulfate solution can be produced by dissolving elemental selenium in sulfite solution at elevated temperature. Precipitation of mercury selenide using selenosulfate reagent was found to be very effective. The precipitation rate proved to be extremely fast, and the formed precipitates have been confirmed to be tiemannite (HgSe) in all experiments. Finally, Solid Waste Disposal Characterization (SWDC) experiments were conducted to examine the mobility of the formed mercury sulfide and mercury selenide. The results show that none of the formed precipitates exceed the Ultimate Treatment Standard (UTS) limit.

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The fundamental dissolution kinetics of silver sulfide and in atmospheric cyanide solutions (2013)

The dissolution kinetics of pure silver sulfide (acanthite) and mercury sulfide (metacinnabar) were investigated using rotating disc and stirred reactor methods and the results were explained using the Levich and shrinking-core Parabolic Leach equations, respectively. It was observed that silver sulfide dissolution was limited by cyanide mass-transport and mercury sulfide by parabolic leaching. Silver sulfide leaching was practically unaffected by pH and dissolved oxygen concentrations while mercury sulfide leaching was sensitive to both parameters. Dissolution rates of both species increased linearly with cyanide concentration and activation energies were calculated using the Arrhenius rate equation as 5.15kJ/mol and 5.81kJ/mol, respectively, which indicates mass-transport control. It was also considered that silver sulfide dissolution is supressed by sulfide-saturation in the solution, while mercury dissolution is inhibited by the growth of sulfur-rich product layers on the particles. Kinetic results were compared to extraction experiments conducted on a spent Heap Leach Residue sample from the Yanacocha Mine in Peru which contained acanthite and cinnabar as its primary silver and mercury species. Extraction kinetics of silver from the Leach Residue was analogous to pure silver sulfide’s experimental results, except that extraction rates decreased with pH. Mercury extraction from the residue was insensitive to both cyanide and pH, but was responsive to oxygen concentration. The discrepancies between pure sample investigations and the Leach Residue suggest that unforeseen interactions with other minerals may be affecting the extraction rates of silver and mercury from the Yanacocha Mine.

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Fundamental aspects of thiosulfate leaching of silver sulfide in the presence of additivies (2012)

Thiosulfate is a promising alternative for leaching silver sulfide ores such as those at theepithermal Yanacocha deposit. These ores suffer from low silver recovery, high mercuryextraction and environmental challenges when treated with conventional cyanide. In thisstudy, cupric-ammonia, ferric-ethylenediaminetetraacetic acid (EDTA), ferric-oxalate andferric-citrate are tested with thiosulfate in a series of rotating disk experiments with silversulfide and with ore from Yanacocha. This thesis publishes experimental evidence thatsupports the use of thiosulfate with different additives as a potential alternative toconventional cyanidation for silver sulfide leaching.The leaching of silver sulfide by cupric-ammonia thiosulfate can occur either by thesubstitution of cupric or cuprous for silver. The cupric catalyzed reaction is favored due to athermodynamic barrier to the cuprous reaction. Rotating disk experiments demonstrate thatcupric-ammonia leaching is under mixed chemical/diffusion control. The leaching rate ismaximized by stabilizing cupric in solution with ammonia and increasing the availability ofthiosulfate for silver dissolution. The addition of EDTA to this system decreased the leachingrate of the silver sulfide disk by lowering the cupric reactivity, but accelerated silver leachingof the ore, likely due to the prevention of passive oxide film formation on sulfides.Ferric complexes used were found to be very unreactive towards thiosulfate, but are reducedby sulfides present in the ore. Ferric-EDTA was the most effective oxidant of the three forleaching silver sulfide with thiosulfate. Silver recovery of the ground ore in batch leachingtests is low due to quartz locking of silver, with cupric-ammonia and ferric-EDTA leaches exposed to air recovering 31% and 26% after 24 hours, respectively. Cyanidation recovered34% silver with a 95% confidence interval of 28-37%. The slightly lower recovery bythiosulfate may be due to silver minerals which are not amenable to thiosulfate leaching.

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Mercury stabilization using thiosulfate and thioselenate (2012)

Mercury is commonly present with gold in nature. As a result it has a tendency to follow gold through the cyanide recovery circuit and ends up in the electro-winning cell as elemental mercury. The laws on the sale and international transport of this mercury are changing. Ultimately, it appears that it will be necessary to stabilize and dispose in a stable form. Mercury sulfide (HgS) and mercury selenide (HgSe) have significantly lower solubilities. The concept of using a thiosulfate dissolution/precipitation method to stabilize mercury as mercury sulfide has been investigated. Comparing the solubilities of mercury sulfide and mercury selenide, mercury selenide is much less soluble. For this reason, the second idea in this thesis is to use sodium thioselenate as a source of selenium in mercury solution to produce mercury selenide.To pursue this project, mercury analysis, mercury leaching and mercury precipitation tests were performed at different temperatures and solution conditions. The resulting solutions were analyzed by Atomic Absorption Spectroscopy (AAS) and the solid precipitates were analyzed by X-ray Diffraction. The EDTA titration method for mercury analysis is effective for a simple mercury nitrate solution. If sodium thiosulfate was added in the solution, thiosulfate interfered with the solution and the titration method was not effective. As a result the AAS method was adopted. Red mercury sulfide can be precipitated by simple aging of mercury thiosulfate solution. Parameters such as temperature, pH and thiosulfate concentration have an effect on the rate and extent of mercury sulfide precipitation. With an increase of temperature, thiosulfate concentration and at lower pH, the mercury precipitation rate increases. However at very high temperature such as 70ºC and 80ºC mercury precipitates as a mixture of red and black mercury sulfide. Thioselenate synthesis was attempted from a mixture of sodium sulfite and selenium powder. The reaction between sulfite and elemental selenium was too slow to be useful.The environmental stability of the mercury sulfide precipitates produced from thiosulfate solutions was investigated. Solid Waste Disposal Characterization (SWDC) tests were done to check the precipitation limit for land disposal and Resource Conservation and Recovery Act (RCRA).

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