Dissertations / Theses on the topic 'Gold leaching'

In this research polyethylenimine coated iron oxide magnetic nanoparticles (PEI-MNPs), as a novel adsorbent, was evaluated for the adsorption of gold from thiosulfate leaching solutions that contain gold, copper and calcium thiosulfate. This novel adsorbent can simply be separated from the solution using an external magnetic field, followed by gold adsorption. Gold elution from the adsorbent was simple and rapid.

Bibliography : pages 157-173.
The main aim of this study was to develop an economically viable bacterial leaching process for a gold-containing pyrite/arsenopyrite ore. The effect of various parameters on, and the mechanism of, bacterial leaching were investigated. Initially milled run-of-mine ore was examined. Batch tests and a continuous bacterial leach were carried out. Bacterial leaching was successful and 91-93% gold dissolution was attained in four days. The process was not economically feasible when compared to the standard flotation-roasting process.

Thesis (MEng (Chemical Engineering))--Cape Peninsula University of Technology, 2019
Over the years, the quantity of e-waste, primarily waste mobile phone printed circuit boards (PCBs), has increased worldwide, as a result of technological advancements in the fields of electronics, telecommunications and computing. Therefore, to sustain the protection of the environment and human health, sustainable measures should be implemented. Electronic waste (e-waste) is known for its wider variety of base and precious metal content compared to naturally occurring ores. Waste mobile phone PCBs are the most attractive type of e-waste due to their higher and more structured precious metal content. Ammonium thiosulphate leaching is considered as one of the promising alternatives to cyanidation in hydrometallurgy. However, its industrial application is limited due to the high reagent consumption which has a direct impact on its cost-effectiveness. This research aimed to investigate the ammonium thiosulphate leaching of gold from waste mobile phone PCBs, with an emphasis on the thiosulphate consumption and interference of the extractable copper in the leaching process. The objectives were to study the effect of acid pre-treatment and copper replacement with nickel as the metal oxidant on gold extraction and thiosulphate consumption. The experiments in this study were all conducted in a batch setup. The PCBs were reduced to less than 3 mm particle size by cutting and crushing. In the acid pre-treatment, sulphuric acid and hydrogen peroxide concentrations were varied between two levels, namely 2 M and 3 M, to identify the optimum conditions that maximised copper extraction and minimised gold extraction. In the ammonium thiosulphate leaching, PCB pre-treatment and metal oxidant were investigated as categorical factors to establish the optimum conditions that maximised gold extraction and minimised thiosulphate consumption. The PCB pre-treatment was varied between two levels: acid-pretreated PCBs and untreated PCBs. The metal oxidant factor was varied between copper and nickel. The other leaching conditions were fixed, as prescribed in the existing literature. The mobile phone PCBs used in this study were found to contain 524 g Au/ton-PCB and 461.8 kg Cu/ton-PCB, as determined by aqua regia leaching. Copper thus contributed to more than 40% of the total PCB mass. In the acid pre-treatment, the optimum reagent combination was 2 M H2SO4 and 3 M H2O2 and resulted in 93.72% copper extraction and 8.83% gold loss in 150 minutes at a pulp density of 50 g/L, 25°C and stirring speed of 350 rpm. Furthermore, it was determined, through material balance, that the PCB mass reduction induced by the acid pre-treatment was mostly attributed to the copper extraction, evidence that the other base metals such as iron and aluminium that could have dissolved in the acid pre-treatment stage were in small quantities in the PCBs to impact the overall mass reduction significantly. The acid pre-treatment was a PCB beneficiation process with an increase in the extractable gold content from 524 to 842 g/ton-PCB and a decrease in the copper content from 461.8 to 51.1 kg/ton-PCB. The variation in H2SO4 concentration had more statistical impact on gold extraction, whereas varying H2O2 concentration had a more statistical influence on copper extraction. The analysis of the goodness of fit of the shrinking-core model to the experimental results indicated that the acid pre-treatment was chemically controlled, with moderate control due to the turbulence. In ammonium thiosulphate leaching, the optimum conditions consisted in using nickel as the oxidant in the thiosulphate leaching of acid-pretreated PCBs at a pulp density of 50g/L, 0.1 M thiosulphate, 0.2 M NH3, 0.03 M Ni2+, pH 10.5, 25°C and stirring speed of 350 rpm. The gold extraction and thiosulphate consumption were found to be 65.41% and 61.03 kg/ton-PCB in 5 hours of leaching time, a significant improvement from the conventional copper-thiosulphate leaching of untreated PCBs which resulted in gold extraction and thiosulphate consumption of 18.61% and 90.9 kg/ton-PCB, respectively. The statistical analysis of experimental results indicated that the variation of the metal oxidant had a higher significance level than the PCB pre-treatment with respect to gold extraction. The assessment of the goodness of fit of the shrinking-core model to the experimental data indicated that the copper-thiosulphate and nickel-thiosulphate leaching processes were both chemically controlled. To assess the degree of comparison between the copper-thiosulphate and nickel-thiosulphate leaching processes for gold extraction from waste PCBs, a preliminary economics assessment was performed, with emphasis on production costs. Considering a basis of 1 metric ton of processed PCB, replacing copper with nickel as the metal oxidant reduced the raw materials costs from 3,768 USD to 2,868 USD, approaching previously reported cyanidation costs closely. The revenue and gross margin were increased from 8,969 and 5,201 USD in copper-thiosulphate leaching to 16,287 and 13,419 USD in nickel-thiosulphate leaching, respectively. Therefore, the use of ammonium thiosulphate as an environment-friendly alternative lixiviant shows potential by virtue of the improved process economics.

The continuous depletion of high grade gold ores has shifted research focus towards the processing of low grade, complex ores containing impurities such as copper which is detrimental to the traditional cyanidation process. Alkaline glycine solution selectively leached copper and gold over impurity elements such as Fe, Ca, Mg and Si from copper – gold ores. Copper was recovered from solution by solvent extraction and gold could be recovered by carbon adsorption while glycine is recycled.

Thesis (Master of Engineering in Chemical Engineering)--Cape Peninsula University of Technology, 2018.
The rate of waste electrical and electronic equipment (WEEE) is growing at an alarming rate, especially in countries where markets are saturated with huge quantities of new electronic goods. Printed circuit boards (PCBs) are a substantial portion of the value contained in waste from WEEE although they are only 6% of the total weight. It is reported that WEEE is currently the fastest growing waste stream in South Africa as the general population’s access to electronic goods in the last decade has increased, especially access to mobile phones. PCBs are found in any piece of electrical or electronic equipment and consist of various metals including precious metals such as gold (Au), silver (Ag) and palladium (Pd). It is reported that gold has the highest economic incentive at 15,200 $ per ton of PCBs. The rapid introduction of new and advanced technology into mobile phones has caused mobile phones to have a relatively short life span, 1 to 2 years to be exact. Mobile phones printed circuit boards (MPPCBs) have more Au content compared to computer circuit boards. They contain 350 g/ton Au whereas computer (PC) PCBs contains 250 g/ton. This research project will recover gold from waste mobile phones PCBs pregnant ammonia thiosulphate leach solution using copper cementation. The cementation process is preferred to all the other technologies of metals extraction from solution due to ultrahigh purity metals that can be obtained and to the less consumption of materials and energy. Electronic parts on the PCBs were manually removed using pliers and screwdrivers. PCBs were then cut to smaller pieces of about 2 x 2 m. The pieces were crushed and milled respectively. Some of the particles were recycled back to the crusher to get finer particles. The particles were separated to particles of sizes between 0 and 1350 μm using a shaker. The comminuted fractions of the PCBs were then used in the leaching step. Batch cementation experiments were performed by bubbling N2 in glass reaction vessel with a working volume of 0.5 L. The reactor was connected to a circulating water bath for temperature control. The recovery percentage of gold at various temperatures, agitation speeds and different amounts of copper powder used, was determined while pH was monitored. The temperature was varied at 30 °C, 40 °C, and 50 °C and the agitation speeds at 300 RPM and 900 RPM. Copper powder was added at 0.5 g/L, 1 g/L, and 1.5 g/L. Gold concentrations were measured by atomic adsorption spectrophotometer (AAS). Scanning electron microscope (SEM) and Energy-dispersive x-ray spectrometry (EDS) analyses of the copper powder after cementation (precipitates) were used to determine the surface morphology and to evaluate the quantitative aspect of the precipitate. It was found that the recovery of gold from ammonia thiosulphate leach solution was greatly affected by agitation speed. At an agitation speed of 900 rpm, 40 °C and 0.5 g of Copper powder, 96% of the gold was recovered from the leach solution. The cementation rate increased as temperature was elevated from 30 to 40 °C, but slightly decreased as the temperature reached 50 °C. The change in experimental conditions affected the gold concentration on the precipitate recovered. This study will provide a possible solution to the WEEE problem and more specifically mobile cell phones, in South Africa.

The focus of this research is the interaction of gold(III)chloride with activated carbon in acidic sulphate solutions. More specifically in this research the in-situ adsorption of gold (III) chloride on activated carbon within an autoclave in the presence of iron sulphide oxidation products, is investigated. Two types of experiments are presented in this research. In the first one, gold(III) chloride (1.5x10-4 M Au(III) plus 0.05 M extra chloride) was allowed to adsorb onto activated carbon (10, 20 and 30 g/L) at pH 1.5 and at different temperatures (22, 60 and 95 oC). Complete gold removal was achieved within a few minutes of contact. The kinetics were found to increase with increasing carbon loading as well as temperature and exhibit "first order" diffusion-controlled behaviour. However an increase in chloride concentration from 0.05 M to 1 M was found to slow down gold adsorption. Via X-ray photoelectron spectroscopic analysis (XPS), it was determined that gold undergoes a reduction upon adsorption on carbon and is present partly as Au(I)Cl and partly as metallic gold. The relative amount of metallic gold increased with temperature but decreased with high chloride ion concentration.The second type (carbon-in-autoclave) of experiment involved pressure leaching (at 155 and 200 oC) of pyrite particulates (acting as model sulphide mineral) spiked with metallic gold (precipitated via cementation) in the presence of low chloride ion concentration and activated carbon. Under these conditions gold was found to lixiviate forming gold chlorocomplexes that reported (in part) on the available activated carbon. The loaded gold, analyzed by XPS was found to be ~2/3 metallic and ~1/3 Au(I)Cl. Because of a "preg-robbing" effect exhibited by hematite forming in-situ, excess sodium was added to the system favouring natrojarosite as the latter has no affinity for gold. In this way gold loading on carbon was maximized. However, significant occurrence of gold deposition on metallic surfaces was observed. By using a glass autoclave the gold plating occurrence was eliminated allowing up to 97% recovery of gold on the activated carbon. Similar results were obtained when copper concentrate material was substituted for pyrite. This work shows that even in small chloride ion concentrations (<0.2 M) gold is easily leached under autoclave oxidation conditions allowing conceptually for its in-situ recovery on activated carbon.
L'objectif de cette recherche est d'analyser le comportement du complexe Au(III)Cl4- en contact avec du charbon actif dans un environnement hydrométallurgique acide contenant des ions de sulfate. Cette recherche détaillera plus précisément l'adsorption du complexe Au(III)Cl4- sur le charbon actif dans un autoclave à température et pression partielle d'oxygène élevée en présence des produits d'oxydation de sulfure de fer. Cette recherche présentera deux types d'expériences. En premier lieu, le chlorure d'or(III) (1.5x10-4 M Au(III) et 0.05 M chlorure) a été adsorbé sur le charbon actif (10, 20, 30 g/L) à un pH de 1.5 et à différentes températures (20, 60 et 95 oC). Le complexe aurifère est complètement adsorbé par le charbon actif après quelques minutes. La cinétique augmente proportionnellement avec une augmentation de la teneur d'or sur le charbon actif de même qu'avec l'augmentation de la température tout en présentant un comportement de diffusion du premier ordre. Toutefois, une augmentation de la concentration de chlorure de 0.05 M à 1 M ralentit l'adsorption d'or sur le charbon actif. Suite à une analyse par rayons-X spectroscopiques photoélectriques (XPS), il a été possible de déterminer que le complexe aurifère (Au(III)Cl4-) subit une réduction sur le charbon actif et est présent en partie comme or natif (métallique) et or complexé Au(I)Cl. La quantité relative d'or métallique adsorbée sur le charbon actif augmente avec la température, mais diminue avec une concentration de chlorure élevée.Le deuxième type (Carbon-in-autoclave) d'expérience consiste à oxyder sous pression (à 155 oC et 200 oC) des particules de pyrite (agissant comme minerais de sulfure idéal) dopé avec de l'or métallique (précipité via cémentation) en présence d'une faible concentration d'ions de chlorure et de charbon actif. Dans cet environnement, l'or est lixivié formant des chlorocomplexes d'or qui ont été adsorbés (en partie) sur le charbon actif. L'or chargé sur le charbon actif a été analysé par XPS et les résultats démontrent la presence de 2/3 métallique et 1/3 Au(I)Cl. Suite au phénomène de « preg-robbing » dû a l'hématite formée durant l`oxydation de la pyrite, un surplus de sodium a été ajouté afin de favoriser la formation du natrojarosite qui n'a aucune affinité avec le chlorure d'or(III). Ainsi, l'adsorption de l'or sur le charbon actif est maximisée. Toutefois, une cémentation importante d'or a été observée sur les surfaces métalliques internes de l'autoclave. En utilisant un autoclave en verre, la quantité d'or plaqué a été éliminé permettant ainsi une récupération de 97% d'or sur le charbon actif. Des résultats similaires ont été obtenus lorsqu'un concentré aurifère de cuivre a été oxydé à la place de la pyrite. Ce travail démontre donc que même de faibles concentrations de chlorure (<0.2M) peuvent facilement lixivié l'or métallique en autoclave permettant théoriquement sa récupération sur du charbon actif durant le procédé d'oxydation.

The kinetics of the dissolution of gold and silver colloids in ammoniacal thiosulfate solutions has been studied using oxygen, copper(II) or oxygenated copper(II) as oxidants at pH 9 - 11 and temperature 22oC to 48oC. The effects of the concentration of the main reagents such as copper(II), ammonia and thiosulfate as well as various background reagents have been investigated. Gold and silver colloids have characteristic absorption peaks at 530 nm and 620 nm respectively. Thus, the extent of gold or silver dissolution in different lixiviant systems was monitored using an ultraviolet-visible spectrophotometer. A comparison of the behaviour of gold colloids and powders has also been made. The beneficial or detrimental effects of silver colloid, and background reagents such as silver nitrate, and sodium salts of nitrate, carbonate, sulfite, sulfate, trithionate, tetrathionate anions have also been investigated. Experimental results show that the relative rates and the extent of gold colloid dissolution at 25ºC in different lixiviant systems in a given time interval are in the order: oxygen-cyanide > copper(II)-ammonia-thiosulfate ≈ oxygen-copper(II)- ammonia-thiosulfate > oxygen ammonia-thiosulfate ≥ oxygen-ammonia > copper(II) ammonia. The analysis of electrode potentials shows that Au(S2O3)23- is the predominant gold(I) species in the lixiviant solutions containing oxygen or copper(II) as oxidant and thiosulfate or mixed ammonia-thiosulfate as ligands. During the reaction of copper(II) with thiosulfate in ammoniacal solution without oxygen, the measured potential using a platinum electrode represent the redox couple Cu(NH3)n2+/Cu(S2O3)m1-2m (n = 4 or 3, m = 3 or 2) depending on the concentrations of thiosulfate and ammonia. The initial dissolution rates of gold colloid by oxygen in copper-free solutions show a reaction order of 0.28 with respect to the concentration of dissolved oxygen, but independent of the concentration of ammonia and thiosulfate. The reaction activation energy of 25 kJ/mol in the temperature range 25°C to 48°C indicated a diffusion controlled reaction. The initial dissolution rates of gold colloid by oxidation with copper(II) in oxygenfree solutions show reaction orders of 0.41, 0.49, 0.60, 0.15 and 0.20 with respect to the concentrations of copper(II), thiosulfate, ammonia, chloride and silver respectively. The presence of silve (I) or chloride ions enhances the rate of gold dissolution, indicating their involvement in the surface reaction, possibly by interfering with or preventing a passivating sulfur rich film on gold surface. An activation energy of 40-50 kJ/mol for the dissolution of gold by oxidation with copper(II) in the temperature range 22°C to 48°C suggests a mixed chemically/diffusion controlled reaction. The dissolution of gold by oxidation with copper(II) in oxygen-free solutions appears to be a result of the reaction between gold, thiosulfate ions and the mixed complex Cu(NH3)p(S2O3)0. The half order reactions support electrochemical mechanisms in some cases. The initial dissolution rates of gold colloid, massive gold and gold-silver alloys by oxygenated copper(II) solutions also suggest a reaction that is first order with respect to copper(II) concentration. High oxygen concentration in solutions has a negative effect on the initial rate of gold dissolution and overall percentage of gold dissolution, indicating that oxygen affects the copper(II), copper(I) or sulfur species which in turn affects the gold dissolution. The surface reaction produces Au(NH3)(S2O3)- and Cu(NH3)p+. The mixed complexes Au(NH3)(S2O3)- and Cu(NH3)p+ re-equilibrate to the more stable complexes Au(S2O3)23- and Cu(S2O3)35- in solution. The dissolution of gold powder by oxidation with copper(II) in oxygen-free solutions shows the same trends as that of gold colloid. The presence of silver(I) or chloride ions enhances the initial rate and percentage dissolution of gold colloid and powder. The dissolution kinetics of gold powder and colloid follow a shrinking sphere kinetic model in solutions of relatively low concentrations of thiosulfate and ammonia, with apparent rate constants being inversely proportional to particle radius. The best system for dissolving gold based on the results of this work is the copper(II)-ammonia-thiosulfate solution in the absence of oxygen or in the presence of oxygen. In the absence of oxygen, copper(II) 1.5-4.5 mM, thiosulfate 20-50 mM, ammonia 120-300 mM and pH 9.3-10 are the best conditions. The presences of carbonate and sulfite have a significant negative effect on the dissolution of gold. The presence of sodium trithionate shows a beneficial effect in the first two hours, while sodium tetrathionate or lead nitrate have a small negative effect and sodium nitrate showed no effect on the dissolution of gold. Silver nitrate and sodium chloride also show beneficial effects. In the presence of oxygen, copper(II) 2.0-3.0 mM, thiosulfate 50 mM, ammonia 240 mM and pH 9.3-9.5 are the best conditions.

Over the last 100 Years, the cyanidation process has been the most popular method for recovering gold from its ores. Despite this, there are still efforts to improve the efficiency of the process, particularly as ores become more difficult to treat. Many investigators have studied the cyanidation process, although a large proportion of these studies have obtained contradictory results. This thesis presents a kinetic and electrochemical study of the leaching of gold in cyanide solutions, and emphasis is placed on rationalising the conflicting results which have been published in the past.The leaching rate of gold was measured using a rotating electrochemical quartz crystal microbalance, an instrument which allows the simultaneous measurement of electrochemical data and mass changes at the solid-solution interface in real time. A proportion of this project was devoted to the on-going design of this instrument, and a number of modifications are discussed in detail. Initially, the leaching of gold in cyanide solutions was investigated under conditions of high purity. Under these conditions, it was found that the gold surface is blocked by a passive film, presumably AuCN. The presence of such a film results in the reaction being chemically controlled, and under typical cyanidation conditions (4 mM cyanide, pH 10.0), the rate of dissolution is very low. These kinetic results were supported by complimentary electrochemical studies, which showed that gold is passive in the potential region where cyanidation occurs.The second part of this thesis presents a study of the effect of system purity on the leaching of gold in cyanide solutions. Solution phase purity was investigated by adding controlled amounts of lead or silver to the leach solutions. It was found that in the presence of low concentrations of lead, the dissolution of gold in 20 mM cyanide solutions was oxygen diffusion controlled (as compared to chemical control for gold in the absence of lead). However, high concentrations of lead were found to be detrimental to the leaching process. It is believed that the role of lead is to modify the surface by cementation, hence reducing the effect of the passive film. Silver was also found to be effective at reducing passivation, and the role of silver believed to be similar to that of lead. It was found that unlike lead, high concentrations of silver are not detrimental to the dissolution of gold in cyanide solutions.Solid phase purity was also found to be important in the leaching of gold, and it was found that the leaching of a gold sample which contains 1 % silver is diffusion controlled. This finding is important from an industrial viewpoint, as most native gold contains some silver. Consequently, attempts were made to rationalise the leaching of gold/silver with current plant practice. Discussion on the effect of cyanide and oxygen concentrations, temperature and lead addition is presented.

Thesis (PhD)--Stellenbosch University, 1992.
ENGLISH ABSTRACT: The dissolution of gold from refractory ores is a complex kinetic problem involving a number of chemical, mass transport and mineralogical factors. In most Witwatersrand ores in South Africa more than 97 % of the gold is dissolved in cyanide medium after a residence time of about 16 hours in pachuca tanks. This high percentage may be the reason why so little fundamental research has been done into the mechanism and kinetics of the leaching process. With the increasingly lower grades of ore mined, the introduction of backfill mining, and the reduction of profit margins, it has become imperative to increase the efficiency of gold dissolution. The effects of the chemistry and particle size on the dissolution of gold in each sample of ore were studied in detail. The emphasis in this study is on the effect of the leaching behaviour of various ore constituents on the rate of gold dissolution. Interferences with the leaching of gold in contact with other minerals or metals could be attributed to the galvanic interaction (electrical conductivity) between the gold and the mineral and to the formation of a surface film on the gold surface. Sulphide minerals and their oxidation products cause the largest decrease in gold dissolution rate. Galena enhances the rate of gold dissolution owing to dissolved Pb(II)-ions. Gold in contact with conductive minerals passivates as a result of the enhanced magnitude of the cathodic cu1Tent. In all experiments the rotating disc of gold passivated so that the rate of dissolution was much slower than that predicted by a mass-transport limiting model. The various films that form on the surface of the gold and associated minerals, as well as the galvanic interaction, depend largely on the pretreatment of the ore. Pre-elimination of host minerals from the gold bearing ore increases the dissolution rate of gold, and explains the kinetics of reaction on the gold surface to a large extent. The selective destruction of the various minerals with oxidative acid leaches destroys and/or decomposes certain minerals which may form films on the gold surface by precipitation. The chemical composition of these films and precipitates depends on the mineralogy of the sample. These films may be oxides, sulphides, carbonates and cyanide complexes. The complexes can be destroyed, depending on the nature of the film, by interstage dilute acid and/or cyanide washes in an agitated vessel. The destruction of the films exposes the gold surface for cyanidation. A simple distribution function similar to the King liberation model is proposed and tested to describe the dissolution step in the multi-step leaching mechanism. For the King model, good agreement is shown with experimental results. For the liberation results obtained by leaching in this study, the trend is co1Tect, but calibration is required for a close fit. A potentially important use for the liberation model by leaching is to predict the leachable or free gold in an ore from the free gold in the complete sample. This approach for studying the leaching behaviour of different gold bearing minerals has provided reasons why some ores leach better than others.
AFRIKAANSE OPSOMMING: Die loging van goud vanuit weerbarstige ertse is 'n komplekse kinetiese probleem wat verskeie faktore soos massa-oordrag, chemiese aspekte en mineralogiese ingeweefdheid insluit. Goud ekstraksies so hoog as 97 % in sianied oplossings in Pachuca reaktore na ongeveer 16 uur logingstyd word behaal in die meeste Witwatersrand ertse in Suid-Afrika. Hierdie hoë ekstraksies mag dalk die rede wees vir die min fundamentele navorsing oor die ekstraksie van goud vanuit minerale in 'n spesifieke erts. Die dalende erts grade, die verlaging van winsgrense en die terugplaas van geloogde erts in die myn noodsaak verbeterde goud ekstraksie. Verkeie faktore nl., chemie, diffusie, partikelgrootte en oplosbaarheid van goud in elke monster erts is in hierdie studie ondersoek. Die sentrale tema was om die logingsgedrag van goud vanuit verskeie minerale in 'n erts te bepaal. Galvaniese interaksie (hou verband met elektriese geleidingsvermoeë) en film vorming is die belangrikste faktore wat die loging van goud in kontak met minerale nadelig beïnvloed. Sulfied minerale en hul oksidasie produkte speel die grootste rol in die verlaging van die tempo van goudloging. Galena verhoog die tempo van goudloging as gevolg van die Pb (II)- ione in oplossing. Goud in kontak met geleidende minerale passiveer as gevolg van die verhoogde katodiese stroomdigtheid. In alle eksperimente met die roterende skyf (goudskyf) apparaat, passiveer die goudskyf in so 'n mate dat die logingstempo baie stadiger is as wat voorspel word met die massa-oordrags model. Die onderskeie films wat vorm op die goud- en geassosieerde minerale se oppervlaktes, asook die galvaniese interaksies, is 'n funksie van die voorafbehandeling van die erts. Die selektiewe eliminering van minerale vanuit 'n gouddraende erts verhoog die tempo van goudloging drasties en dit beskryf die kinetika van goudloging op die goudoppervlak in 'n groot mate. Die selektiewe eliminering van minerale deur gebruik te maak van oksiderende suurlogings, vernietig sekere van die minerale wat films op die goudoppervlakte kan veroorsaak deur middel van presipitasie. Die chemiese samestelling van hierdie films hang af van die mineralogie van die monster. Dit bestaan meestal uit oksiedes, sulfiedes, karbonate en sianiedkomplekse en hulle kan vernietig word deur middel van inter-stadia verdunde suur-en/of sianied wasse. 'n Eenvoudige distribusiefunksie, soortgelyk aan die King bevrydingsmodel word voorgestel en eksperimenteel getoets om die logingstap in die multi-stadia logingsmeganisme te beskryf. Vir bevryding deur loging, is die neiging van King se model korrek, maar kalibrasie word benodig vir goeie passing. 'n Potensiele gebruik van die aangepaste model is om vrye of loogbare goud in 'n spesi fieke partikel grootte fraksie van 'n erts te voorspel as die vry goud in die totale fraksie bekend is. Die resultate uit hierdie studie kan gebruik word om die logingsgedrag van gouddraende minerale te voorspel en te beskryf, en dus veduidelik hoekom goud uit sekere ertse beter loog as uit ander.

Although bacterial leaching of sulphidic minerals is a well-known phenomenon, it is only in the last ten years that full-scale bacterial leaching plants have been commissioned for gold processing. In order for bacterial leaching to compete successfully with other pretreatment processes for refractory ores, particularly with established technologies such as roasting and pressure leaching, it needs to be efficient. This requires the optimization of the parameters affecting the leaching reaction and the growth of bacteria. The entire biotreatment process is agitation leaching, carried out in stirred reactors or Pachuca type reactors. The bacterial oxidation is a complex reaction involving gaseous, liquid and solid phases. The interactions are highly complex, and analysis is complicated by the presence of solids in the leaching medium. Inspite of the amount of research that has been performed, kinetic and process models are underdeveloped. Since kinetic data varies widely with the type and source of concentrate, experimental data should be generated before doing the full-scale reactor design. In sizing reactors for a commercial scale process, it would be useful to have a mathematical model that one could use to predict the amount and rate of release of metal, as a function of the various operating parameters of the system. G.R.Halli arsenical gold sulphide concentrate obtained from Hutti Gold Mines Ltd., Karnataka, was chosen for our study, because of its high refractoriness. An indegenous strain of Thiobacillus ferrooxidans was used for biooxidation. The experiments were conducted in a well-agitated stirred tank reactor under controlled conditions. Sparged air was supplemented with carbon-dioxide for optimized growth. In this work, more than 90% gold and 95% silver could be recovered from the sulphidic gold concentrate when bioleaching was used ahead of cyanidation, compared to 40% and 50% by direct cyanidation. A generalized model, which accounts for both direct bacterial attack and indirect chemical leaching, has been proposed for the biooxidation of refractory gold concentrates. The bacterial balance, therefore, accounts for its growth both on solid substrate and in solution, and for the attachment to and detachment from the surface. The overall process is considered to consist of several sub-processes, each of which can be described in terms of a mechanism and related rate expressions. These sub-processes were studied seperately under kinetically controlled conditions. The key parameters appearing in the rate equations were evaluated using the experimental data. Since the refractory concentrate contains pyrite and arsenopyrite as the major leachable entities, leaching studies have been done on pure pyrite and arsenopyrite as test minerals and the key parameters in the rate equations are evaluated using this data. The model so developed is tested with the leaching kinetics of the concentrate. The growth of bacteria is dependent on the availability of the substrate, ferrous iron, and the dependence is modelled by the widely accepted Monod equation. The effect of carbon dioxide supplementation on the bacterial activity was studied and the optimal concentration for growth was found to be l%(v/v). Studies on indirect chemical leaching showed that the rate is sensitive to surface area of concentrate. Indirect rate constant of arsenopyrite was found to be greater than that of pyrite, since pyrite is more nobler than arsenopyrite. Conditions of direct leaching alone was obtained at high pulp density and using substrate adapted bacteria. The rate constant of arsenopyrite was found to be greater than that of pyrite. The parameters obtained were tested with the overall batch leaching data of the concentrate and favourable comparision was obtained. Thus, it has been possible to isolate the various simultaneous sub-processes occurring during the leaching and propose useful models to describe these processes in some detail. The model has been extended successfully to predict the continuous leaching behaviour using the parameters obtained from the batch data. Studies on the effect of residence time and pulp density on steady state behaviour showed that there is a critical residence time and pulp density below which washout conditions occur. The critical residence time at 10% pulp density was found to be 11 hrs. Operation at pulp densities lower than 5% and residence times lower than 72 hrs is not favourable for efficient leaching. Studies on the effect of initial ferric iron concentration showed that there exists an optimum concentration of ferric iron at which the time required to reach steady state is minimum.

Over the last 100 Years, the cyanidation process has been the most popular method for recovering gold from its ores. Despite this, there are still efforts to improve the efficiency of the process, particularly as ores become more difficult to treat. Many investigators have studied the cyanidation process, although a large proportion of these studies have obtained contradictory results. This thesis presents a kinetic and electrochemical study of the leaching of gold in cyanide solutions, and emphasis is placed on rationalising the conflicting results which have been published in the past.The leaching rate of gold was measured using a rotating electrochemical quartz crystal microbalance, an instrument which allows the simultaneous measurement of electrochemical data and mass changes at the solid-solution interface in real time. A proportion of this project was devoted to the on-going design of this instrument, and a number of modifications are discussed in detail. Initially, the leaching of gold in cyanide solutions was investigated under conditions of high purity. Under these conditions, it was found that the gold surface is blocked by a passive film, presumably AuCN. The presence of such a film results in the reaction being chemically controlled, and under typical cyanidation conditions (4 mM cyanide, pH 10.0), the rate of dissolution is very low. These kinetic results were supported by complimentary electrochemical studies, which showed that gold is passive in the potential region where cyanidation occurs.The second part of this thesis presents a study of the effect of system purity on the leaching of gold in cyanide solutions. Solution phase purity was investigated by adding controlled amounts of lead or silver to the leach solutions. It was found that in the presence of low concentrations of lead, the dissolution of gold in 20 mM cyanide solutions was oxygen ++
diffusion controlled (as compared to chemical control for gold in the absence of lead). However, high concentrations of lead were found to be detrimental to the leaching process. It is believed that the role of lead is to modify the surface by cementation, hence reducing the effect of the passive film. Silver was also found to be effective at reducing passivation, and the role of silver believed to be similar to that of lead. It was found that unlike lead, high concentrations of silver are not detrimental to the dissolution of gold in cyanide solutions.Solid phase purity was also found to be important in the leaching of gold, and it was found that the leaching of a gold sample which contains 1 % silver is diffusion controlled. This finding is important from an industrial viewpoint, as most native gold contains some silver. Consequently, attempts were made to rationalise the leaching of gold/silver with current plant practice. Discussion on the effect of cyanide and oxygen concentrations, temperature and lead addition is presented.

Includes bibliographical references.
Auriferous conglomerates of the Archaean Witwatersrand Basin in South Africa host one of the largest known gold resources and rate as the world’s most outstanding example of a fossil megaplacer deposit. For the past 40 years, Witwatersrand gold production in South Africa has been progressively declining due to problems related to high energy costs, decreasing grade, accessibility to greater depths, health and safety issues, labour union unrest and economic uncertainties: thus the overall viability of current gold production is questionable. Ultimately, the future of Witwatersrand gold mining relies on devising smarter strategies across the entire industry, but in particular critical areas such as comminution and extraction. With the continuous increase in mining depth, dominance of low-grade gold ores and strict safety regulations, metallurgical processing options have become limited. Heap leaching is a well-established technology which continues to grow in use and provides several benefits to solve some of these problems. High pressure grinding rolls (HPGR) is another technology with significant potential, especially for its application in coarse particle heap leaching due to its ability to induce micro-cracks as well as its high grinding efficiency and low energy requirements. This study explores the use of these two technologies in a process mineralogical framework using novel 3D X-ray computed tomography mineralogical analysis in order to assess a potential of the Witwatersrand gold ore for heap leaching.

Bench and pilot scale testing has shown that biooxidation of a refractory arsenical sulphide concentrate enhanced gold and silver extraction by cyanidation. Direct cyanidation of Southern Tail bulk sulphide concentrate typically resulted in less than 10% gold and silver recovery. Biological pretreatment of the sulphide concentrate resulted in 70% gold extraction and at least 30% silver extraction within economic process constraints. Biooxidation of a low grade arsenical sulphide concentrate, 5.5 g Au/t, 80 g Ag/t, and 2-5% As, was economically attractive due to apparent preferential oxidation of the gold bearing suphide mineral assemblage(s). Selective sulphide oxidation minimized the bioleach residence time, the oxygen/CO₂ requirements, and the costs associated with neutralization of the acidic byproducts of sulphide oxidation. Approximately 15% total sulphide oxidation was required to effect 70% gold extraction and 30-40% silver extraction. Additional sulphide oxidation improved gold extraction slightly and improved silver extraction significantly but was not economically attractive. Complete batch and continuous laboratory bioleach facilities were constructed at Equity. A detailed description of apparatus, methods, and operating data is presented. Critical variables that were studied included temperature, pH, pulp density, particle size, air/CO₂ requirements, nutrients, bioleachate recycle, inoculum source, degree of sulphide oxidation, cyanidation conditions, and bioleachate treatment. The laboratory data provided the design and operating criteria for a continuous bioleach pilot plant. Operation of the continuous pilot circuit, with a nominal design capacity of 2 tonnes per day, showed there was no significant effect of scale on the biohydrometallurgy or resultant precious metals hydrometallurgy. A detailed description of the pilot equipment and steady state operating parameters is presented. Considerations for scaleup and for plant scale design and optimization are discussed. A preliminary feasibility study, based on a conceptual design for an 800 tonne per day bioleach circuit, is detailed. The feasibility analysis showed that it is economically attractive to scavenge gold and silver values from Southern Tail flotation tailing using biooxidation to enhance the gold and silver extraction by cyanidation. The base case cash flow analysis showed a Net Present Value potential of $4,975 x 106 (1985) at 15% discount factor.
Applied Science, Faculty of
Mining Engineering, Keevil Institute of
Graduate

L'objectif de l'étude est l'application d'un procédé basé sur la lixiviation statique suivie de l'électrolyse directe (LSED) a la valorisation de petits gisements d'or, notamment les petits gisements amazoniens du brésil dont la production représente 75% de l'or produit dans ce pays, bien que compte tenu des techniques artisanales utilisées les taux de récupération de l'or ne dépassent pas 30 à 40% comme le montre l'état de l'art. Le procédé LSED a été appliqué à 3 minerais types différents quant aux métaux accompagnateurs et à la para genèse: il s'agit du minerai du chapeau de fer du gisement du rouez (France), du filon de quartz d'essakane (Burkina Faso) et du minerai sulfure complexe de volta grande (Brésil). Les taux de mise en solution varient de 70 à 90%. L'électrolyse directe effectuée sur les solutions de lixiviation diluées (3 à 8 mg au/l) permet d'atteindre des rendements d'électrodéposition très élevés, mais alors les rendements faradiques sont très bas et par conséquent les consommations énergétiques très élevées. L'analyse des dépôts électrolytiques au microscope électronique à balayage montre des structures non orientées, plurimétalliques et de morphologie variable. Le procédé LSED peut s'appliquer industriellement a un minerai d'or pauvre en métaux accompagnateurs, avec lixiviation en tas de 2000 t et en employant des électrolyseurs d'un volume unitaire de 1000 l minimum. Dans ce cas, après environ 7 h d'électrolyse directe, on peut récupérer 60% de l'or avec une consommation d'environ 800 kwh/kgau

The kinetics and mechanism of gold leaching in the ammonia-thiosulfate-copper (ATS-Cu) system have been studied using the rotating electrochemical quartz crystal microbalance (REQCM). Anodic polarization, cathodic polarization and leaching experiments were included in this study. The effect of concentration of different reagents, applied potential, pH, temperature, and electrode rotating velocity on the gold leaching rate have been investigated. The effect of solution copper species on the anodic reaction of gold dissolution in thiosulfate solution was elucidated using the REQCM. With respect to the role of copper on the anodic processes, two possible mechanisms were proposed. It was shown that, in the absence of S₂O₃²⁻, the electrochemical reaction on the cathode is the reduction of Cu(NH₃)₄²⁺+ to Cu(NH₃)₂⁺ in the potential range of 0.2 to -0.3 V vs. SHE. Based on other experimental results, it is believed that, in the presence of S₂O₃²⁻, the electrochemical reaction on the cathode also is the same as the reaction in the absence of S₂O₃²⁻. It was found that the anodic process and leaching process are under chemical reaction control, but the cathodic process is under diffusion control. It may therefore be concluded that the leaching process is under anodic control under the majority of conditions tested (cathodic control will be important when the cupric ammine species is very dilute in solution). Experiments show that increasing the ratio of NH3/ S₂O₃²⁻ will favor the leaching rate. However, excess thiosulfate or excess ammonia may inhibit the dissolution of gold. The effect of selected additives was also studied.

To develop a fundamental understanding of cyanide heap leaching of low-grade gold ores, with significant copper mineralization, two sets of column leach tests were operated under two different leaching solutions. In one set, four 1.5 m tall column tests were leached with sodium cyanide solution. In the second set, another four columns were leached with solutions of cyanocuprate complexes in the absence of free cyanide. Each set was run for a period of up to 200 days. The behavior of individual species, were investigated for the purpose of developing reliable rate parameters for the mathematical model. Significant amount of weak acid soluble copper minerals was discharged from the ore, simply by pre-rinsing with water. The cyanide leaching is shown to occur in different stages with remarkable characteristics as follows: (a) Significant decrease in the pH at the beginning of the cyanide column leach tests; (b) Reductive dissolution of the remaining WAS copper salts and dissolution of copper sulfide minerals, from the first day of the experiment; (c) Dissolution of gold from the first stage while no free cyanide is present in the solution; (d) The occurrence of two distinct peaks in copper concentration, first as the result of leaching of labile copper sulfides and second as the result of redissolution of copper cyanide precipitates; (e) Free cyanide breakthrough in the effluent, which is accompanied by the appearance of dissolved iron. The second set of column leach tests showed that the leaching of gold with a solution of cyanocuprate complex and in the absence of free cyanide is accessible with rate parameters comparable to leaching with cyanide. It is also concluded that some of the copper minerals can be leached in the absence of free cyanide. Iron does not show any dissolution in cyanocuprate leaching systems.

In recent years, emissions of hydrogen cyanide from metallurgical operations have received renewed attention by legislative bodies, leading to the need for a reliable quantification method for HCN volatilisation. Subsequently, the purpose of this project, launched by Anglogold Ashanti Ltd. and in collaboration with MINTEK and the University of Pretoria, was to develop a prediction model for cyanide volatilisation from plant operations and tailings storage facilities in South Africa. The study was done in four stages, the first being a laboratory study of the equilibrium behaviour of hydrogen cyanide. Henry’s Law constant (kH) was determined at different solution cyanide concentrations, salinities and temperatures. A value for kH was established at 0.082 atm.L/mol, which was found to be independent on the solution cyanide concentration between 10 and 200 ppm cyanide. In addition, the effect of temperature on kH was found to be negligible at solution temperatures between 20 and 35ºC. It was also concluded that high salinities increase kH and promote volatilisation, but this effect was negligible at the typical salinity levels found in South African process water. The second stage entailed a detailed study of the mass transfer coefficient, KOL, for hydrogen cyanide from cyanide solutions and pulp mixtures, both in the laboratory and on-site. It followed from this investigation that the most important parameters affecting KOL are the HCN (aq) concentration in the liquid, the wind velocity across the solution or pulp surface, expressed in terms of a Roughness Reynolds number, Re*, and the moisture content, or solid to liquid ratio, of the pulp. Furthermore, it was concluded that KOL is highly sensitive to HCN (aq) concentrations at low concentrations, while it becomes rather insensitive to HCN (aq) at concentrations above 20 ppm HCN (aq) . The data generated by the laboratory and on-site test work was incorporated into the development of an empirical prediction model, based on the Roughness Reynolds number (Re*), moisture content (M), and aqueous cyanide concentration (HCN(aq) ) which may be described by the following equation: KOL= a Re*b Mc HCN(aq)d + e The model coefficients were subsequently determined for application of the model to leach tanks, adsorption tanks, tailing storage facility surfaces and return water dams. The calculated model predictions for KOL were in excellent agreement with the measured test work data. Finally, the prediction model was validated at the leach and adsorption sections of a selected gold plant and a selected tailings storage facility. The model predicted that 9% of the cyanide lost in the leach and adsorption section could be attributed to HCN volatilisation. As for the tailings storage facility, the model assigned 63% of the cyanide lost from the tailings storage facility to HCN volatilisation, of which 95% occurred from the area on the tailings dam surface covered in a thin liquid film. It is recommended that the current methods available for the determination of HCN (aq) be further improved, due to the sensitivity of the model to the input value of the HCN (aq) concentration, in order to ensure that reliable predictions are made. It is also suggested that additional validation work be done in order to establish the generic applicability of the model to different sites.
Dissertation (MEng(Metallurgical))--University of Pretoria, 2007.
Materials Science and Metallurgical Engineering
unrestricted

A mathematical model of cyanidation heap leaching of cupriferous gold ores which includes the speciation of all cyanide, cyanate, thiocyanate, hydroxide and sulfate complexes, and their corresponding copper and gold precipitates, was developed to facilitate the treatment of these complex ores. The concepts of back mixing and solution retention are represented by two fundamentally different fluid elements: i) the bulk solution flowing through the heap, and ii) a stagnant solution reservoir adjoining and enclosed by the ore particles. Four 1.5 m tall column tests simulating actual heap leaching behavior under different nominal influent cyanide concentrations were operated for a period of up to seven months, along with semi-batch coarse ore and batch fine ore leaching tests. The resulting experimental data were used for the calibration and validation of the model. It was found that, in the presence of significant copper sulfate mineralization, the pH of the solution would naturally buffer to approximately 4.6. Further pH decrease to 3.5 is provoked by the oxidation of cyanide to cyanate by cupric ions. At such low pH levels, it is thought that cuprous and aurous cyanide salts tend to precipitate out of solution both within and around porous ore particles. As a result, neither dissolved gold nor dissolved copper from secondary copper sulfides appear in the column effluent for a significant period of time, depending on the influent cyanide concentration and the height of the column. Once dissolved again, gold appears in the column effluent, whereas its leaching kinetics are a strong function of leaching kinetics of secondary copper sulfide minerals such as covellite and chalcocite. Furthermore, the combined effect of covellite leaching and copper sulfate discharge is synergistic in that both promote the precipitation of copper hydroxide. This leads to the dispersal of a Cu(OH)₂(s) reserve throughout the column which has the effect of increasing the long term cyanide consumption. This, in turn, affects the leaching of secondary copper sulfides, and hence, of gold.

The application of ammonium thiosulfate for the treatment of copper-gold ores has been investigated. Leaching studies were conducted with copper minerals, copper minerals with gold addition to solution and copper-gold samples of different copper and gold grades. The behaviour of thiosulfate, tetrathionate and sulfate in solution was studied using ion chromatography. The copper sulfide minerals chalcopyrite and enargite seem to be unreactive toward an ammonium thiosulfate leach. Covellite and chalcocite leach to a slight extent in this leaching system. The copper extractions of the sulfide minerals seem to be independent of the availability of complexing agents. The copper oxide minerals cuprite and malachite showed high copper extractions in the presence of sufficient lixiviant. Experiments showed that both gold extraction and thiosulfate stability are influenced by a combination of aeration and cupric ions in solution. High initial gold extractions were achieved in an aerated solution in the presence of cupric ions. However, these conditions simultaneously catalyzed thiosulfate degradation, resulting in gold precipitation. Therefore, it is important to establish a balance between providing sufficient air and cupric ions for fast gold dissolution, and to minimize the amount of air in the presence of cupric ions to prevent excessive thiosulfate degradation. A promising potential alternative to these conditions is a 24 hour leach without forced aeration.

The effects of some associated minerals on thiosulfate gold leaching were studied through thermodynamic analysis and leaching experiments on composite ore samples containing various minerals and a reference silicate gold ore. In the leaching test on the reference gold ore, about 93% of gold was extracted within 3.0 hours. The presence of various amount of pyrite, pyrrhotite, chalcopyrite, arsenopyrite, chalcocite, bornite, and some lead species, has significant detrimental effects. Under reduced oxygen conditions, the thiosulfate consumptions could be significantly reduced. High gold extractions (i.e. >= 90%) were observed in the leaching tests with reduced dissolved oxygen (i.e., 0.7% oxygen in the supplied gas) in the absence or in the presence of sulfide minerals such as pyrite, pyrrhotite, arsenopyrite and chalcopyrite. High copper concentration and a pre-aeration step was also found to largely increase the gold extractions under such conditions. Thiosulfate-copper-ethylenediamine system was found effective in the leaching of gold. The leaching kinetics was significantly slower than that of the conventional thiosulfate-copper-ammonia leaching. The consumption of thiosulfate, however, was largely reduced. This leaching system worked effectively on the reference gold ore within a wider pH range (e.g., 6-11), with or without ammonia. The presence of ammonia in a low concentration improved the leaching rate but also increase the consumption of thiosulfate. Comparable gold extractions were observed in the leaching of the composite ores containing various sulfide minerals, such as pyrite, pyrrhotite, chalcocite, galena and chalcopyrite. The leaching of gold in the presence of iron sulfides was also improved by applying chemical additives, such as, carbonate, calcium, galena, phosphate, and additional hydroxide anion. It is proposed that these additives either passivated the harmful surface of sulfide minerals or masked some detrimental aqueous species. Finally, some improved leaching methods concluded in this study were applied on a few industrial ore samples in order to demonstrate the effectiveness of these methods. It was found that by comprehensively applying these improved thiosulfate leaching strategies, satisfactory gold extractions and thiosulfate consumption results were archived on these ores.
Thesis (Ph.D, Mining Engineering) -- Queen's University, 2008-09-18 11:48:38.672

Thiosulfate has shown considerable promise as an alternative to cyanide for gold leaching. However, one of the main limitations of the thiosulfate system is the high consumption of thiosulfate. Besides increasing the cost of the process, the degradation products of thiosulfate have been claimed to passivate gold surfaces and the polythionates often produced are loaded onto resins proposed for gold recovery. The thiosulfate degradation process is not completely understood. Of the degradation products, trithionate is a concern in the resin recovery of gold and is persistent in gold leach solutions. Very little is known about the expected behaviour of trithionate, both with respect to its formation and its interaction with other solution species. The focus of this work was therefore to further the understanding of the behaviour of trithionate in gold leach solutions. Experimental work was carried out to determine the kinetics of trithionate degradation in systems resembling gold leaching solutions, and a kinetic model was derived for trithionate degradation. The rate of degradation of trithionate in aqueous ammoniacal solutions was expressed by Equation 1. [the equation is not included] In some cases, the presence of lower concentrations of thiosulfate catalysed the reaction while excess thiosulfate inhibited it. However, under typical gold leaching conditions, thiosulfate was not expected to have a significant effect so was excluded from Equation 1. Cupric copper was not found to have any significant effect on the rate of trithionate degradation under the conditions tested. This observed trithionate degradation rate equation was integrated with known kinetic behaviour of thiosulfate and tetrathionate based on literature findings to develop an overall model for the thiosulfate degradation and the resulting solution speciation of the sulfur oxyanions in the absence of ores. The model was evaluated against experimental data and its shortcomings were identified. The model parameters were adjusted to obtain a best fit to the experimental data. It was found that the best-fit parameters varied with the experimental conditions, indicating inadequacies in the model. The main concern was that the understanding of the thiosulfate degradation reactions is limited, and the way in which thiosulfate degradation was described had a major impact on the model output. In particular, the effects of copper species and pH on thiosulfate degradation have not been adequately addressed in the literature. Even taking into consideration the limitations of the model, based on the model output, decreasing the cupric concentration and increasing the ammonia concentration should help to minimise thiosulfate degradation. Solution recycle can also be used to minimise thiosulfate degradation but can result in a build up of trithionate. Limiting the reaction time would also be useful. This investigation has led to an improved understanding of the behaviour of trithionate in gold leach solutions and the model of the thiosulfate degradation system is a first step in developing a useful assessment method for thiosulfate degradation and solution speciation under gold leaching conditions. Further research is required to refine the model, particularly with respect to thiosulfate degradation to trithionate and tetrathionate.
Applied Science, Faculty of
Materials Engineering, Department of
Graduate

Biooxidation is an attractive process for unlocking of gold from refractory ores, but could be further improved by reducing the rather high cyanide consumption during the cyanidation of the biooxidation residues. The high cyanide demand is typically ascribed to the presence of metastable intermediate sulfur species formed during the oxidation of the sulfide, but it is also conceivable that jarosite formation during the acidic biooxidation process and subsequent leaching during the alkaline cyanidation could also account for significant cyanide consumption. The present work was done to establish if jarosite would readily form under typical biooxidation conditions and, if so, if it would then leach during cyanidation. Methods to stabilise or remove the jarosite were also investigated. To this end crystalline potassium jarosite was synthesized simulating conditions for the mesophile BIOX? process and its decomposition in alkaline media at pH 10.5 to 11 at 25?C investigated. It was found that it transformed into another amorphous iron compound as indicated by a change in colour from yellow to reddish, the presence of potassium and sulfate ions in solution, as well as the disappearance of the characteristic XRD pattern of jarosite. The jarosite also leached in aqueous alkaline cyanide as indicated by the presence of iron in solution, with increased leaching at higher cyanide concentrations. The jarosite could not be totally passified by aging, but removing it by leaching with an iron-complexing agent like oxalic acid was found to be possible. The consumption of cyanide by jarosite during gold leaching could thus be significant, but would obviously depend on how much jarosite is formed during the bioleaching process. Jarosite was observed to be present in the mesophile and thermophile BIOX? products from Fairview Mine with more jarosite present in the products of the thermophile process. The process jarosite was significantly more stable than that produced in the laboratory during alkaline cyanidation with 17 mass percent of the jarosite converted to ferric hydroxide for the process jarosite compared to complete conversion for the laboratory jarosite for a 24 hours leaching period. However, it still accounted for a significant amount of the cyanide consumed.
Dissertation (MSc)--University of Pretoria, 2016.
tm2016
Materials Science and Metallurgical Engineering
MSc
Unrestricted

This project consisted of a literature and a laboratory study to investigate the applicability of an electrochemical model, to the leaching kinetics of gold ore. The procedure was to get an approximate measurement of the gold surface potential and then, to use only this potential and an area term in a two parameter rate expression. The potential measurements in a pulp, with a gold and gold-silver electrode, were attempted but, the electrodes became passive. The results showed that the rate controlling mechanism was a combination of the rate of both oxygen and cyanide diffusion and it was concluded that the applicability of an electrochemical model to the leaching kinetics of gold ore, could not be ruled out. It was found that the leaching kinetics could be modelled very accurately by using the solution redox potential, however, no theoretical basis could be found to explain this observation.

A dissertation submitted to the faculty of Engineering and the Built Environment, University of Witwatersrand, in fulfillment of the requirements for the degree of Master of Science in Engineering, 2020
The cyanidation process of refractory gold ores has been the subject of numerous investigations aimed at improving the gold recovery and the leaching kinetics. Many pretreatments methods have been developed in this regard and numerous new leaching configurations have been introduced in the process. However, even after many improvements, the leaching of refractory gold ore remains a process of some duration (24 hours or more to reach acceptable gold recoveries). From recent literature investigations, hydrodynamic cavitation has been found to be a promising new approach which may advantageously enhance the cyanidation process. This approach results in the enhancement of mass transfer kinetics of multiphase streams due to impacting two pulp streams against one another in a vessel called the “Jetleach reactor (Jetleach)’’. In this work the Jetleach reactor was applied to three different refractory gold ores. Two rougher flotation concentrates from West Africa (DIO and MVO) and one flotation concentrate from the Ergo plant in South Africa. Results from acomparative cyanidation test between the Jetleach process and conventional methods, here represented by the normal leaching in agitated vessel and bottle rolling methods, have shown impressive results. On the DIO sample, the Jetleach reactor has shown an improvement of almost 8% and 19% gold recovery compared to the normal leaching and bottle rolling method, respectively, while on the MVO sample, the improvement was about 10% compared to the normal leaching method and 17% compared to the bottle rolling tests. On the Ergo sample, the improvement of gold recovery was over 12% compared to both two conventional methods. Cyanidation using one factor at a time (OFAT) experiment revealed that the pump pressure, the feed pulp solid percentage and the oxygen flowrate are the parameters significantly influencing the gold recovery. However, cyanidation, using design of experiments (DOE) and response surface methodology strategy, has shown that, in addition to these three significant parameters, three other interactions between parameters also have a significant effect on the gold recovery. The analysis done on the Jetleach residues, using the Scanning Electron Microscope (SEM), have revealed the presence of cavities and cracks on the surface on some particles which could explain one reason for the faster leaching kinetics and higher gold recoveries. The particle size analysis, undertaken on the feed and the residue of the Jetleach reactor, revealed a size reduction of particles after passing through the reactor. This size reduction can also be one of the elements contributing to faster leaching kinetics. In pursuing the understanding of the generation of cavitation in the Jetleach reactor, a visualization test using a high-speed camera was conducted on a transparent tube reactor, operating with water to simulate the process in the Jetleach. Images collected from the high-speed camera have shown the presence of cavities generated in the water and the collapsing process of same. These images also revealed that cavitation was directly connected to the pump pressure. The higher the pressure of the pump, the larger the resulting cavitation zone. This work has shown that the Jetleach reactor is able to improve the leaching kinetics and the gold recovery of all three refractory gold ores involved in this project. The reactor was also able to reduce the cyanide and oxygen consumption on the Ergo sample. Therefore, an integration of this reactor in the cyanidation process, will result in benefits in terms of both metal recovery improvement and reagent consumption reduction
CK2021

M.Tech.
The efficiency of hydrochloric acid leaching of gold from low grade gold-bearing material has been investigated in the presence of sodium hypochlorite (NaOCl) to produce in situ chlorine gas which is an oxidizing agent able to oxidize gold metal to gold soluble forms: Au+ and Au3+. The effect of concentrations of HCl, NaOCl, and mixtures HCl + NaOCl was sought. An investigation on chlorine species was conducted to predict their stability areas. The reactions involved and their mechanisms were established. Prior to leaching, the feed was subjected to fire assay, Atomic absorption spectroscopy (AAS), X-ray diffraction (XRD), X-ray fluorescence (XRF), and Fourier transform infrared spectroscopy (FTIR) analyses, to quantify the various elements, and to find out the chemical composition and the different mineralogical phases which are present. The main minerals found in the feed were: quartz, pyrite, muscovite-2M1, (M); and clinochlore. The grade was found to be 0.62 g/t. It was found from the hydrochloric acid leaching process that most of metals constituents (iron and potassium) of the feed were leached and consumed all the hydrochloric acid. Therefore, leaching of gold from low grade gold-bearing materials in aqueous chlorine solution is not an effective process for now. Hydrochloric acid leaching of cobalt and copper from four oxidized cobalt-bearing ores was studied in the presence of FeCl2 to produce in situ Fe2+ that is a reducing agent able to reduce Co3+ to Co2+ which is the soluble form of cobalt. The effects of the concentrations of HCl, FeCl2, mixture HCl + FeCl2, leaching time, particle size, and the dissolution kinetics were investigated. Prior to leaching the feeds were characterized with XRD, XRF, FTIR, AAS, Scanning electron microscopy (SEM), and gravimetric analysis. Cobalt contents were found to be between 2.59% and 39.76% in the four ores namely, high Cu ore, high Co ore, low Co ore, and high Co ore with mica. Effect of FeCl2 concentrations in HCl solutions involving 35.06 g; 70.13 g; and 105.18 g corresponding to 0.25 M; 0.5 M; and 0.75 M respectively, were studied and the reactions mechanism established. It was found that an increase in FeCl2 concentration increases the Co dissolution due to the reduction of Co3+ into Co2+ which is soluble. High Co extractions of 99%, 99%, 99%, and 95% were obtained when mixtures 1 M HCl + 105.18 g FeCl2 were used as lixiviants for high Cu ore, high Co ore, low Co ore, and high cobalt ore with mica, respectively. Copper recoveries of 88%, 77%, 75%, and 86% were obtained for high Cu ore, high Co ore, low Co ore, and high Co ore with mica, respectively. The increase in FeCl2 concentration did not have any important effect on the Cu extraction. Mechanisms of involved reactions were established. The order of cobalt leaching reaction was found to be 2 with respect to cobalt concentration. Optimum leaching conditions were found to be: leaching time: 30 minutes, reagent concentrations: mixture 1 M HCl + 105.18 g FeCl2, S/L ratio = 1:20, stirring speed avoiding the vortex, room temperature, pH values from 2.28 to 0.03, and ORP values from 0.402 to 0.322 V.

M.Tech.
The study presented, investigated the acid leaching behavior of calcines emanating from conventional roasting and microwave assisted roasting of sulphide ores. The aim was to investigate a possible change in the mineral composition of the calcine produced and to investigate the leaching behavior of the calcine produced by conventional and microwave assisted roasting. Two sulphide concentrates were received, pyrite concentrate from OTJIHASE concentrator and sphalerite concentrate from ZINCOR. The mineralogical study was done using the XRD (PANalytical X-PertPro X-Ray diffractometer) to determine the mineral composition of the concentrate and the head grade of each concentrate was determined with an Atomic Absorption Spectrometry (AAS) machine for copper, iron and zinc. The concentrates were wet screened at 500, 355, 212, 150, 106 and 75 μm screens for size distribution and for the purpose of the study it was required that 80% of concentrate should pass the 75 μm screen since after roasting, the calcine obtained were to be acid leached. Conventional roasting was carried out at temperatures of 500oC, 700oC, 850oC and 1000oC at durations of 5min, 10min, 20min, 30min, 1hour and 2hours, while microwave assisted roasting carried out at powers of 200W, 400W, 600W, 800W and time intervals of 3s, 10s, 30s, 30min and 1hour

Thesis (MTech. degree in Metallurgic Engineering)--Tshwane University of Technology, 2008.
The purpose of this study is to develop a feasible process for extracting gold and silver from an electrolytic zinc refinery residue. Composite samples of electrolytic zinc refining lead-silver residues containing traces of gold (1.35 g/t) and significant amounts of silver (155 g/t) were subjected to batchwise cyanide leaching to assess the feasibility of extracting gold and silver. This project was concluded with a recommendation that a pilot Residue Leach Plant should be commissioned to beneficiate gold and silver from the residue dam for a better estimation of lime, cyanide consumptions and gold and silver recoveries. The return on capital investment will be high at current gold price and operational cost is low. It is also recommended to investigate the economics of cyanide leaching of a pre-concentrate obtained from froth flotation.

Complementary electrochemical and spectroscopic techniques were used to characterize the behavior and composition of the passive layer formed at the gold-thiosulfate interface in the presence of copper. Raman studies of three different cationic (calcium, ammonium and sodium) thiosulfate leaching solutions showed that the concentrations of sulfate, thiosulfate, trithionate and tetrathionate remained constant. Initial leaching current densities for the three systems were identical, however, significant differences were noted in the open circuit potentials of these systems. Gold nanorod electrodes were employed as substrates for Surface Enhanced Raman Spectroscopy (SERS) studies of the gold-thiosulfate interface. The composition and behavior of the passive layer at the gold- thiosulfate interface greatly differed from that of the bulk solutions. Higher order polythionate species were not observed, and significant differences were noted in the behavior of species common between the three thiosulfate leaching solutions. Passivation levels determined from SERS indicate that in the presence of copper, the cation associated with thiosulfate may play a key role in the extent of passivation on the gold surface.
Barrick Gold Corp., NSERC

A thesis submitted to the Faculty of Engineering and The Built Environment, University of the Witwatersrand, Johannesburg, in fulfilment of the requirements for the degree of Doctor of Philosophy in Engineering Johannesburg 12 September 2016
Comminution and leaching unit processes play a major role in extracting valuable minerals from ore. Most of the research reported in the literature has focused on optimising individual unit operations rather than on integrating the whole process. This thesis develops an integrated approach to mineral processing systems and flow sheets and is intended to create a methodology for process synthesis that can be applied throughout the extractive metallurgical industry. This could lead to improved efficiency in the overall process by obtaining optimum recovery and, most important, a reduction in energy and material costs. In order to illustrate the methodology a particular example was chosen, namely optimizing the joint comminution and leaching of a particular gold ore. In this investigation laboratory scale grinding and leaching profiles for a gold feed sample (1700–850 μm) were measured. In a laboratory mill various combinations of grinding media, filling level and ball size were investigated, and of the three ball sizes used (10, 20 and 30mm) breakage was most pronounced for the 20 mm. Thus for instance it was also established that when using a higher filling ( =30%) and a ball size of 30 mm, more energy was consumed but less liberation occurred, thus a lower amount of gold was extracted during a 24-hour leaching period. Finally, the breakage kinetics of the gold ore was looked at. Using a standard population model the breakage and selection function parameters were successfully calculated. An investigation into the dissolution kinetics of gold ore in a solution of NaCN was also done. These were found to depend on the stirring rate, reaction temperature, particle diameter and the concentration of the leachant. The rate increased with the stirring speed, reaction temperature and leachant concentration, but decreased when the particle size was greater. The activation energy for the dissolution was estimated at about 3 kcal/mol. Furthermore, the linear relationship between the rate constant and the reciprocal of the square of the particle size is a strong indication that the gold dissolution process is diffusion-controlled. The experimental results were well-fitted to a shrinking core model. In attempting to understand the results, the researcher carried out a number of experiments that involved an investigation into the relationship between comminution and leaching in terms of energy usage and particle size, the former to establish the most efficient application of energy, and the latter to identify the degree of fineness that would ensure optimal recovery. The Attainable Region (AR) method was then used to establish ways of finding the leaching and milling times required to achieve minimum cost (maximise profit). No work on utilizing the AR technique to minimise the cost of milling and leaching on a real industrial ore has previously been published. The investigation aims to show how the AR technique can be used to develop ways of optimising an industrial process that includes milling and leaching. The experimental results were used to show how this method could be successfully applied to identifying opportunities for higher efficiency when performing these operations. The approach however is general and could in principle be used for any two or more unit operations in determining how the product from one unit should be prepared to feed to the next unit so as to optimize the overall process.
MT2017