David Dreisinger

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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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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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 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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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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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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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 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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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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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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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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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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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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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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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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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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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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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 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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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 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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