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The role of sulfur species in the passivation of gold surface in cyanide and thiosulfate solutions was investigated using density functional theory (DFT). The interaction energy between various sulfur complexes as well as deposited sulfur allotropes with the gold surface were computed and the contribution of each specie to the formation of a passivated layer were compared. For this purpose, a new parameter was introduced to explain and understand the tendency of each sulfur complex to interact with the gold surface from a molecular point of view. It was observed that the ratio of the hydration to the binding energy (H/B) of each sulfur specie can be well correlated with the calculated interaction energy with the gold surface. This parameter could be effectively used to predict the interaction of each specie with a substance, only based on the structural characteristics of the specie. It has been also demonstrated that the theoretically calculated energies of interactions of sulfur complexes with the gold surface are well correlated with experimentally reported data. The results showed that S, S2−, polysulfides, tetrathionate, and sulfite species affect the gold leaching process directly through the passivation of the gold surface. Sulfate, dithionate, trithionate, and thiocyanate do not contribute to the passivation effect, although they indirectly affect the gold leaching process by consuming the oxygen and cyanide required to dissolve gold in the cyanidation process. The sulfur allotropes forming on the gold surface also passivate the gold surface and limit the access of the lixiviants required for the gold leaching process. The results could shed light on the mechanism of passivation of gold surface by sulfur species in gold leaching and this effect can be controlled by inhibiting the formation of main sulfur contributors in passivation process.

Due to the drawbacks of copper ions used to activate pyrite flotation in gold processing plants where free cyanide occurs in process water and the benefits of lead ions in gold cyanidation, this study investigated the counteraction of lead ions on the negative effects of free cyanide on pyrite flotation in an alkaline condition. Flotation results indicate that 50 ppm (6.8 × 10-5 mol/L) free cyanide sufficiently depressed pyrite flotation, while 9.5 × 10-4 mol/L lead ions substantially activated pyrite flotation. It was interesting to find that 9.5 × 10-4 mol/L lead ions significantly counteracted the depression of 50 ppm free cyanide on pyrite flotation but did not completely restore the flotation of pyrite activated by the same amount of lead ions. Cryogenic X-ray photoelectron spectroscopy was then applied to determine the surface species responsible for pyrite flotation in different conditions. Iron cyanide and thiocyanate formed on pyrite surface were attributed to the depression of free cyanide on pyrite flotation, while lead species, in particular lead hydroxides, adsorbed on pyrite surface were attributed to the activation of lead ions on pyrite flotation. The adsorption of lead ions on pyrite surface was enhanced by the adsorbed cyanide species and lead ions counteracted free cyanide in depressing pyrite flotation through the covering of lead species on the top of iron cyanide and thiocyanate. However, lead species preferentially adsorbed and accumulated on the top of thiocyanate, limiting the capability of lead ions in counteracting free cyanide.

A description of the agitation leach plant used for treating gold-copper ores at Anglo Asian Mining’s Gedabek mine in Azerbaijan is given. The plant employs a unique resin-in-pulp system for gold recovery. The integration of the agitation leach plant into the existing heap leach/ADR/SART operation is described. Operational problems caused by increasing amounts of copper in the Gedabek gold ores are outlined. Previous work on the use of ammonia in copper-gold cyanide leach systems is reviewed. The test work leading to the introduction of ammonia in the industrial leaching plant is described. In the plant it was found that the extraction of copper was halved by the use of ammonia, thus leading to significant reductions in the amount of cyanide consumed in the full-scale leaching system.

The re-treatment of gold tailing heaps is a potentially significant revenue stream for both larger and smaller scale gold companies. Limited access to capital and lack of skilled labour remain restrictive especially when dealing with a refractory ore and/or where limited amount of material are available. Alternative techniques such as low temperature pre-leaching options to unlock value are therefore in demand. Several low temperature pre-leaching experiments in acidic and alkaline solutions, followed by cyanidation were performed which ultimately led to the further investigation of a caustic pre-leaching step. The Au recovery were found to be a balance between the Au being released for cyanidation through gangue mineral dissolution and the Au being leached/lost in the pre-leaching step. Depending on the NaOH concentration, the major elements (excluding Na) in the pre-leach solution were found to be Si (500–8500 mg/L), S (2400–1800 mg/L), Al (350–50 mg/L), Zn (10–100 mg/L), Fe (60–75 mg/L) and As (20–55 mg/L). Au recovery in the subsequent cyanide and elution step increased with pre-treatment temperature and NaOH concentration, reaching a peak before either stabilising or decreasing. Longer pre-leaching residence time and a lower solids percentage resulted in less Au being recovered.

A novel approach for the management of solid and liquid wastes of mining was implemented in this work. Unmodified and modified metallurgical solid wastes, sulfate tailings, were used as adsorbents for the removal of cyanide ions from synthetic mine process water. Energy dispersive X-ray fluorescence (ED-XRF), scanning electron microscopy (SEM) and nitrogen adsorption methods, based on Brunauer-Emmett-Teller (BET) theory, were used for analysis of composition, structure and surface area of the adsorbents. A wet-granulation method was used for preparation of granules from sorbent powders with polyvinyl acetate as a binder and the further modification of the sorbent surface was carried out by atomic layer deposition (ALD) of Al 2 O 3 . The granules were used for the removal of cyanide ions from synthetic solutions by batch adsorption method. ALD-modified sorbent showed approximately 3.5 mmol g −1 adsorption capacity for cyanide ions. The adsorption isotherm was fitted with Langmuir and Freundlich adsorption models. The cyanide ions on the modified sorbent surface is enabled to form a stable and non-toxic cyanide complex in the solution, which can be reused as the complexing agent in the metallurgical industries.

It is known since the end of the XIX century that bromine is an excellent leaching reagent for gold but the research concerning its potential application as an alternative reagent has been relatively despised, mainly because elemental bromine is hazardous and extremely difficult to store or transport in safe conditions. We performed experimental tests in order to assess the feasibility of bromine produced in-situ, in the leaching reactor, to dissolve gold using ore samples from the unexploited Portuguese mine of Castromil. The reagents used in the production of bromine (NaBr, NaOCl and HCl) can be more safely stored than bromine. The ore sample was crushed and grinded in order to produce a particle size distribution (D80 = 0.48 mm) that would guarantee that most of the gold and electrum would be liberated in accordance with previous mineralogical studies. The following operating conditions were assessed: temperature, solid/liquid ratio, composition of the leaching reagent, stirring speed, residence time, effect of pretreatments and the addition of oxidants to the bromine solution. The highest gold dissolution (89%) was achieved with a previous roasting, although the simultaneous addition of H2O2 and FeCl3 allows for a similar recovery (80%) with much lower environmental impact. These results were achieved with the following leaching solution: 4.4 g NaBr, 30 mL NaOCl, 42 mL HCl and 28 mL H2O and with the following operating conditions: a solid/liquid ratio of 1, temperature around 95 °C, stirring speed of 450 rpm, natural reaction pH and a leaching time of 4 h. The comparison between bromine, cyanide and thiosulfate leaching for the same ore indicates that bromine is a good technical alternative for leaching this gold ore. The following advantages may be stressed: fast dissolution rate, non-toxic reagents, it does not require the use of pH regulator and can be performed with low quantities of leaching solution.

Galvanic interactions between different types of grinding media and arsenopyrite and pyrite, during the grinding was quantified and the effect of galvanic interactions on gold cyanide leaching was investigated. The synthetic ores were prepared by mixing arsenopyrite or pyrite with clean gravity gold concentrate and quartz. In the presence of arsenopyrite, the results showed that when the grinding of the ore was performed using the 30% chromium or ceramic grinding media, the gold leaching was more efficient than that when the forged steel media was used. Based upon the measurement, it was suggested that the galvanic interactions between the forged steel grinding media and sulfide minerals resulted in the formation of more iron hydroxide than that when the 30% chromium or ceramic grinding media was used. As a result, free cyanide was converted to ferrocyanide thereby lowering the cyanide available for the gold leaching. Additionally, the galvanic interactions also reduced the dissolved oxygen content and thus cyanide gold leaching was more difficult particularly at the early stages of the leaching process because the dissolved oxygen is required to dissolve gold. In contrast, when the grinding was performed with the 30% chromium or ceramic media, a very small amount of free cyanide was converted into by-products. Cyanide gold leaching was the most efficient when the 30% chromium media was used during the grinding of the ore for both pyrite and arsenopyrite. In addition, in the presence of pyrite, due to the lower galvanic current between pyrite and grinding media in comparison with arsenopyrite, the amount of iron hydroxide was lower, and thus more free cyanide was available for the gold extraction. These results indicated that more gold was extracted in the presence of pyrite rather than the arsenopyrite for a given media type used. More precisely, in the presence of pyrite, when the 21% chromium grinding media was used, 70% of gold was extracted while it was 60% in the case of arsenopyrite in 24 h. The selection of grinding media appears to be significant when cyanide reagent consumption is evaluated.

Precipitation of jarosite-type compounds (jarosites) is widely used in hydrometallurgical processes to control metal impurities in solution. During this process, economically valuable elements like silver may co-precipitate. Due to the toxicity of cyanide, used for silver leaching, alternative complexing agents are needed. These agents should have low human and environmental toxicity and efficiently extract silver from ores. In this study, thiosulfate (S2O32−) and thiocyanate (SCN−) were used as alternatives to cyanide in silver extraction from jarosites. The silver extraction and leaching rates obtained with these leaching agents were compared with those obtained with cyanide. Sodium jarosite containing silver was synthesized by the hydrothermal method and characterized by different analytical techniques. Experimental data obtained from batch leaching experiments were fitted with the chemical reaction-controlled shrinking core kinetic model based on studies reported in the literature for similar systems. The effects of pH, temperature, concentration, and the type of leaching agent on the leaching rates were studied. It was found that, under the same temperature and pH conditions, the best extraction efficiency of silver with S2O32− and cyanide were 99.6 % and 97.2%, respectively, but there was a noticeable higher consumption of reagent with S2O32−. The best silver extraction with SCN− was only 35.4 %. However, the leaching rates and the kinetic parameters obtained were very similar, indicating that the kinetics of the jarosite dissolution and the silver complexation could be similar, independent of the complexing agent used. The global kinetic models obtained for extraction of silver from jarosites are useful for evaluating the viability of new complexing agents.

The ongoing depletion of gold and silver-containing ore deposits demands a shift towards the processing of more complex sources. Leaching of gold or silver from these sources, which are named refractory and are typically low-grade, can be complicated due to one or more of the following effects: (1) kinetic or thermodynamic impediments on cyanide diffusion or on the dissolution of minerals, (2) readsorption of leached gold or silver (preg-robbing), and (3) excessive cyanide consumption due to the presence of cyanicidal agents. This article reviews the occurrence of these effects as a function of the elemental composition and compiles strategies to overcome them. The first type of refractory behavior was found to be related to arsenic or antimony sulfides and sulfosalts, manganese oxides, permanganates, iron sulfides, oxides, and hydroxides, tellurides, as well as different sulfur-containing minerals. In general, most kinetic or thermodynamic impediments are addressed by performing intensive grinding and through oxidation processes for sulfides, such as roasting, pressure oxidation, chemical oxidation or biological oxidation. On the other hand, preg-robbing activity was described as common in ores that contain carbonaceous matter, silicates, and sulfides. Strategies to deal with preg-robbing ores often include the destruction or deactivation of the preg-robbing surface, or the addition of another adsorbent to compete with the preg-robber (carbon-in-leach). Finally, cyanicidal agents include sulfides and compounds containing base metals, mainly copper, iron, zinc, and nickel. In the case of sulfides, performing oxidation processes is common, while base metals often require one or more pre-leaching stages in which they can be partially removed from the ore prior to cyanidation. Even though this review intends to serve as a compendium of strategies directed at improving cyanidation processes depending on the nature of the elements present in a feed, a final discussion stressing the importance of developing ore-specific strategies is given.

The main purpose of this research has been to evaluate and optimize the application of hydrodynamic cavitation (HC), combined with hydrogen peroxide, as a promising process for the effective degradation of cyanide in aqueous effluents. The experimental work was carried out using cavitation equipment with a venturi device connected to a tubular circuit which allowed a closed-cycle flow to run for 120 min, in which the effect of control parameters as inlet pressure, H2O2:CN─ ratio, pH, and temperature have been evaluated for the treatment of solutions with initial cyanide concentration in range 100 to 550 mg L─1. The results showed that in optimal conditions cyanide degradation using only HC reached 70% and, using solely H2O2 as oxidizing agent it reached 63%. Efficiency of the combined treatment process was evaluated on the basis of their synergetic effect as it turned out to be more effective showing a 99.9 % cyanide degradation in less than 120 min. The optimum set of conditions that produced the highest degradation rate and efficiency was: inlet pressure 4 bar; pH 9.5; and H2O2:CN─ ratio = 1.5:1. The process was also evaluated on the basis of cavitational yield and in terms of energy and chemical treatment costs. The results have demonstrated that the combined treatment technology of HC + H2O2 can be effectively used as a fast and highly efficient treatment of wastewater containing cyanide.

A mixture of green organic compounds (MOC) with powerful depression to both pyrite and pyrrhotite was introduced to replace NaCN in the separation of highly complicated Zn-S bulk concentrate. The results of single, batch as well as pilot flotation tests indicated its excellent selective depression in the Zn-S bulk concentrate flotation. Especially, the comparison of the pilot tests showed an evident improvement in overall concentrate grade as well as Zn recovery with the mixed depressant MOC replacing NaCN, and the corresponding values were 3.03% and 5.01%, respectively.

Carbonaceous materials found in gold ores often exhibits strong preg-robbing properties which adversely affect gold recovery during the cyanidation due to its ability to adsorb, or preg-rob gold from the cyanide leach solution. Treatments of carbonaceous ores with chemical surfactants as blinding agents to limit preg-robbing have shown promising results however, if not properly used, could have a strong detrimental effect on the gold recovery process. The present study investigated the effectiveness of using commercially available surfactants (Dextrin, naphthalene sulphonate and Triton X-100) as blanking agents to minimize the preg-robbing capacity of various carbonaceous materials (activated carbon, anthracite and bituminous) and carbonaceous gold ore through adsorption and leaching studies. Chemical analysis of ore sample used for the study showed that the gold (Au) content was 17.2 g/t. The organic carbon content of the ore sample was 0.95%. Mineralogical composition of the ore sample determined with X-ray diffraction (XRD) showed that pyrite (24.8%), Quartz (21.5%) and Muscovite (21.8%) were the main mineral in the ore. Gold adsorption studies performed revealed that the treatment of the carbonaceous materials (activated carbon, anthracite and bituminous) with the blanking reagents significantly decrease gold adsorption by the carbonaceous materials. However, the blanking ability for the surfactant varied for the various carbonaceous materials. Application adsorption studies to gold leaching test confirmed that the three surfactants, especially naphthalene sulphonate could be used as blanking agents for carbonaceous gold ores during cyanidation. Gold recovery from CIL studies increased from 46.0% to 67.7%, 52.3% and 49.0% after treatment with naphthalene sulphonate, dextrin and triton X-100 (49.0%) respectively.

In this paper, we proposed a combined leaching-electrowinning approach to recover Cu from AVR sediment. Firstly, speciation diagrams of the metal-cyanide (Cu, Zn and Fe) system were constructed using thermodynamic calculations. Afterward, leaching-electrowinning experiments were systematically conducted. The speciation calculations showed that CuSCN was converted into Cu-CN compounds when the CNT/Cu mole ratio exceeded 1.71. The cyanide ions complexed with Zn2+ were completely captured by Cu+ when the Cu/Zn mole ratio reached 5.0, yielding the formation of zinc hexacyanoferrate composite or Zn(OH)2 precipitation. Corresponding leaching results indicated that the selective Cu leaching from AVR sediment was realized with the leaching efficiency of 99%, forming Na2Zn3[Fe(CN)6]2 as the main ingredient of the residue at the optimal leaching conditions. Cyclic voltammetry tests suggested that cuprocyanide complexes were dominantly presented as [Cu(CN)2]− complexes in the leachate (5680 mg/L, CNT/Cu mole ratio of 2.18), exhibiting much higher current density than that of [Cu(CN)3]2− in the cyanide effluent. However, during the electrowinning process, [Cu(CN)2]− were gradually converted into higher-order complexes, and the cathodic current efficiency decreased correspondingly. Ultimately, the Cu recovery of 92% was achieved with a current efficiency of 57% and energy consumption of 1.91 kW·h/kg Cu.

Copper ions are often added in the flotation of gold-bearing pyrite to activate pyrite so that high gold and pyrite recoveries can be achieved. However, the residual copper ions are problematic in subsequent cyanidation by consuming cyanide. There is a great need to replace copper ions by lead ions which are beneficial to cyanidation. This study compared lead and copper activation on pyrite with different degrees of surface oxidation at pH 5.0. Flotation tests indicated that lead and copper ions were equally efficient in activating slightly oxidised pyrite and both produced a high pyrite recovery. For moderately oxidised pyrite, copper ions were superior to lead ions in activating pyrite. However, for severely oxidised pyrite, lead ions prevailed over copper ions in activating pyrite. By Cryogenic X-ray photoelectron spectroscopy analysis, it was found that the different behaviour of lead and copper ions in activating pyrite was governed by the lead and copper species adsorbed on pyrite surface. In general, pyrite oxidation generated iron and sulphur oxidation species which inhibited the adsorption of lead and copper ions on pyrite surface. Lead ions had stronger affinity to oxidised pyrite than copper ions but pyrite flotation required a higher amount of lead-activating products than copper-activating products.

In this work, the detailed analysis of HCN volatilization, taking place in tailing storage facilities, was made. Volatilization experiments were performed at conditions typical of gold leaching industrial tailing ponds. The meticulous statistical analysis (including full factorial 33 experimental design) let to determine the variables and their interactions affecting the percentage of HCN volatilization. Volatilization tests were performed in an open, temperature-controlled, continuously-stirred batch reactor. The percentage of HCN volatilization was directly proportional to the temperature and temperature-pH interaction and inversely proportional to the pH, cyanide concentration, and pH-pH and temperature-cyanide concentration interactions. HCN volatilization was promoted at acidic conditions. A first order rate law was used to represent the volatilization rate. The specific rate constant (k) was found to be following function of temperature and pH: k ( T,pH ) = A 0 · exp - 18760.78 T , where: ln ( A 0 ) = ( 0.11 ± 0.11 ) · pH + ( 58.08 ± 0.16 ) . The obtained kinetic model represented properly (R2 = 0.90) experimental data in a wide range of industrial conditions: cyanide concentration (300–2000 mg·L−1), pH (3–9), and temperature (16–20 °C). The increase in temperature, from 16 to 20 °C, let to the increase in k, by a factor of ca. 2.5 ± 0.8. The increase in solution pH, from 3 to 9, provoked its decrease, by a factor of ca. 1.9 ± 0.3.

It is known that the presence of sulphidic minerals in the cyanidation of gold ores may cause significant consumption of oxygen supplied in the injected air. This may result in dissolved oxygen starvation for the oxidative leaching of the gold, and ultimately it will reduce the maximum attainable recovery of gold from the ore. In addition, the presence of sulphides leads to extra consumption of the cyanide-leaching agent, NaCN, due to the formation of thiocyanate, therefore increasing costs. These types of gold sulphidic ores may be pre-treated prior to cyanidation by means of an oxidation step, converting the sulphides into oxides or sulphates. This treatment leads to a reduction in the consumption of dissolved oxygen and of cyanide in the cyanidation step and to an improvement in the metallurgical recovery. In the current work we present the results of a five month full-scale trial carried out in a gold extraction plant in Brazil, which normally operated with three tanks in series carrying out an alkaline pre-oxidation step using compressed air only, followed by a train of fourteen aerated and mechanically-agitated tanks for the cyanidation. The ore feeding the leaching circuit averages 1.70 g Au/t, with about 2.5% of pyrrhotite (FeS) as the main sulphide constituent. The addition of 60 L/h of concentrated hydrogen peroxide, H2O2, 50% w/w (density = 1.19 g/mL) for pre-oxidation of a slurry of 60% solids at a rate of 150 t/h (dry ore) resulted in a marked increase in dissolved oxygen (DO). This addition corresponds to a dosing rate of 0.24 kg 100% H2O2 per ton of dry ore and increased the dissolved oxygen level from an average of about 1.0 to 7.2 mg/L in the pre-oxidation tanks. It also led to an overall reduction of NaCN consumption from an average of 0.52 to 0.40 kg/t of ore, and an increase in metallurgical Au recovery from an average of 91.3% to 92.5%.

Many aspects of gold ores leaching by glycine (C2H5NO2) and glycine-cyanide solutions are still not well understood. The base-metal components have been shown to have a strong impact on gold leaching kinetics. In this research, with the aim of reducing cyanide consumption, leaching of a polymetal (Pb-Cu-Zn) gold ore using a synergistic lixiviant mixture of cyanide and glycine has been studied. The results showed that by increasing cyanide concentration from 500 to 1500 ppm, the gold dissolution increased from 46% to about 80%. The gold dissolution of 90% could be achieved by adding 0.5–1 M/L glycine to 500 ppm cyanide while the copper dissolution increases to more than 30%. Interestingly, 80% gold dissolution could be obtained by 0.5 M/L glycine even using 200 ppm cyanide over 72 h. At the optimal approach, cyanide consumption can be reduced by about 80% compared with 1500 ppm cyanide (without glycine). In the studied sample, besides Cu, Pb, Zn and Fe minerals are the other cyanide consumers. Their dissolutions increased as glycine concentration increased. This can be considered as a negative aspect of the cyanide-glycine leaching and indicates its poor selectivity for gold. Pb, Zn, Fe, Ag, and Cu are more sensitive to glycine concentration rather than cyanide concentration.

Thiocyanate is among the most promising alternative gold lixiviants for gold leaching and has not been extensively covered in the literature. Thiocyanate is 1000 times less toxic than cyanide and is a very strong gold leaching lixiviant. This review paper provides an overview of the chemistry, thermodynamics, kinetics, and processing of gold leaching using thiocyanate as well as gold recovery from the pregnant leach solution. One main reason that the thiocyanate gold leaching process has not yet been commercialized is that the process requires a reduction-oxidation (redox) potential higher than that required of cyanide gold leaching. The oxidant, mostly ferric, is needed to leach gold. However, the redox potential must be lowered by removing oxidants to allow gold recovery from the pregnant leach solution. The molar ratio of oxidant to thiocyanate is considered an effective parameter for gold oxidation kinetics and should range between 2 and 20. Thiocyanate gold leaching is an electrochemical reaction which occurs at pH 1.5–2.5 and electrochemical potential of 600–700 mV. The literature shows fast gold leaching kinetics in well-designed thiocyanate systems. The solubilized gold-thiocyanate species can be recovered from the pregnant leach solution by a variety of methods including conventional activated carbon adsorption, solvent extraction, ion exchange, and cementation. Leaching can be performed in both heaps and tanks.

Pyrite is a gangue mineral widely associated with valuable minerals. It is problematic to deal with in differential flotation because it easily reports to flotation concentrates. Pyrite may float due to natural or self-induced hydrophobicity or collector-induced hydrophobicity. Its flotation can also be enhanced by the activation of copper or lead ions emanating from other minerals. Depressants are generally required to selectively depress pyrite. In this article, a comprehensive review of the fundamental studies of depression mechanisms and effectiveness of commonly used reagent systems to depress pyrite is presented. It includes inorganic reagents (hydroxyl ions, oxygen conditioning, cyanide, sulfoxyl reagents, etc.), organic reagents such as polysaccharides (starch, dextrin, carboxymethyl cellulose etc.), polyacrylamides, wood-extracted biopolymers, diethylenetriamine and a combination of these methods. Overall, the action of pyrite depressants is either to desorb the collector or activator from pyrite surface, deactivate the activating ions, prevent collector adsorption on pyrite, or make the pyrite surface hydrophilic. The mechanisms of some non-xanthate collectors such as dithiophosphate, dithiophosphinate and thionocarbamate in rejecting pyrite were also briefly reviewed.

The adsorption of cyanide ions (CN−) on pyrite, marcasite and pyrrhotite surfaces was studied using density functional theory. The depressing effects of CN− on these three minerals were then investigated by flotation tests. The calculated results show that the adsorption energy of CN− on the marcasite surface is the largest, −374.91 kJ/mol; followed by pyrrhotite, −364.34 kJ/mol; and pyrite, −327.94 kJ/mol. These results indicate that the interaction between CN− and marcasite surface is the strongest, then pyrrhotite, followed by pyrite. The depression effect of cyanide on the pyrite, marcasite and pyrrhotite in concentrate with increasing NaCN dosage was analysed by X-ray diffraction (XRD). The results show that the depression effect is in decreasing order of marcasite, pyrrhotite and pyrite, and the greater the cyanide dosage, the stronger depression effect is, which is in good agreement with the calculated results.

Flotation performance of gold-bearing pyrite in refractory gold ores depend upon the interplay among process water chemistry, grinding chemistry, and ore mineralogy. The nature of pyrite itself (finely disseminated reactive pyrite vs. low reactivity) also plays a role in its surface reactivity. This paper reviews the experimental studies examining the interactions between grinding media, gold bearing pyrites and process water occurring during grinding of refractory gold ores and the effect of these interactions on flotation response of pyrite. The literature review reveals that the interactions are highly convoluted. It appears that the galvanic interactions between forged steel media and sulfide minerals ores are particularly significant after grinding of massive sulfide ores. Process water chemistry can also have an amplifying effect, as the process water becomes more laden with cyanide species and other ionic species. This paper should contribute an improved understanding of electrochemical and chemical processes occurring during the grinding of refractory gold ores, which is essential to improve flotation performance of pyrite. Judgement is reserved with regards to the overall economics associated with the use of various grinding media. Media consumption due to wear will have to be balanced against flotation recovery of gold bearing sulfides and whole circuit behaviour where gold in tailing may also be further recovered through leach processes.

Recently, leaching of gold with cyanide has come under scrutiny for its possible detrimental effects to the environment. Also, in some gold bearing ores and concentrates, conventional cyanide leaching is not the best method for gold extraction. This is due to gold particle encapsulation, cyanicides or pregrobbing carbon such as found in refractory ores and concentrates. Hence, there is much interest in developing an alternative way to extract gold from ore. Possible benefits of leaching gold with something other than cyanide include easier processing of refractory ore bodies including: carbonaceous, cyanicides, and sulfides, cheaper reagent and operating costs and opening of mine properties in places where cyanide use has been limited by the law. Hence, this paper details the fundamentals and development of the Alkaline Sulfide Gold Leaching system.

Cyanide is the most commonly used depressant to reject pyrite as a gangue mineral at alkaline pH in differential flotation. However, the depression mechanism of cyanide still remains uncertain. The depression effect has been proposed as the preferential adsorption of cyano species on pyrite, which has, however, not been confirmed experimentally. It is also believed that cyanide inhibits the chemisorption of thiol collectors due to the decrease of surface redox potential. In order to clarify the controversial depression mechanisms, the interactions between pyrite and cyanide at pH 10 were investigated in this study. Preferential adsorption of cyanide on pyrite rather than the decrease of surface redox potential was determined to be the depression contributor. Surface cyanoferric species were identified, both electrochemically and spectroscopically, on pyrite surface in the presence of cyanide, inhibiting the oxidation of pyrite and the oxidation/chemisorption of xanthate. This study provides new insights into the depression of pyrite flotation by cyanide.

The organic–inorganic hybrid polymer Al(OH) 3 –polyacrylamide (Al–PAM) was investigated in flocculation of particles in 10 wt% cyanide tailing suspensions in this study. The effect of Al–PAM was evaluated with photoelectric turbidimeter, focused beam reflectance measurement (FBRM) and particle video microscope (PVM). It was found that as the amount of Al 45 –PAM with maximum zeta potential and molecular weight increased, the initial settling rate of particles increased, reaching the highest 11.7 m/h, 18.5 times of the blank without flocculants. Correspondingly, the turbidity of the supernatant was 0.95 NTU, compared to 291.6 NTU in the blank, which was over 98% of turbidity removal efficiency. The cationic Al–PAM demonstrated better flocculation to the particles in the cyanide tailing suspensions than the commercial non-ionic guar gum (GG), and has potential for commercialization .The electrostatic repulsion between tailing particles was regarded as a critical parameter in determining polymer flocculation and our study indicated that both charge neutralization and bridge adsorption contributed to the particle flocculation.

In the present paper, influence of a lignin-based biopolymer in improving gold and silver recoveries during cyanidation leaching is discussed. Series of laboratory cyanidation leaching testworks at the absences and the presences of the biopolymer were performed under variations of the biopolymer concentration, NaCN concentration, slurry density and particle size distribution. The ore samples used were from Pongkor Gold Mine of PT. Antam, Tbk. in Indonesia. The biopolymer additive was found to reduce slurry viscosity that leads to a higher dissolved oxygen concentration in the leach solution. The testwork results show that the use of biopolymer improves gold and silver recoveries as well as leaching kinetics. At biopolymer dosage of 800 mg/L, the increases of gold and silver recoveries were 4.5% and 6.4%, respectively. The lignin-based biopolymer is able to disperse fine solid particles and stabilizes the ore suspension in the leach-slurry that results in a better contact between the ore and the leaching agents. The effect of biopolymer on the gold and silver recoveries was found to be more significant at lower cyanide dosages. The addition of biopolymer resulted in the increase of cyanide consumption, lowering final pH and higher dissolved iron which is associated with gold-bearing pyrite dissolution. Enhanced dissolution of gold-bearing pyrite at higher dissolved oxygen concentration was found as the major factor that improves gold recovery at the presences of the biopolymer in addition to lowering slurry viscosity that leads to higher dissolved oxygen concentration and effect of fine particles dispersion.

In gold processing plants, the process water contains a considerable amount of weak acid dissociable (WAD) cyanide species. The cyanide-containing water recycled to flotation circuits has a deleterious effect on mineral flotation. Zinc cyanide can be a major constituent of the process water due to the cyanide leaching of zinc minerals and the zinc cementation applied to precipitate gold and silver. In the present study, the effect of zinc cyanide on the flotation of gold from a pyritic ore was evaluated and the competitive adsorption between zinc cyanide species and the collector, potassium amyl xanthate (PAX) or Aerophine promoter (3418A), on gold was studied by surface enhanced Raman spectroscopy (SERS). It was found that free cyanide did not depress the flotation of liberated gold, but zinc cyanide with a CN/Zn ratio of 2 did. The depression of gold flotation by zinc cyanide was due to the preferential adsorption of Zn(CN) 2 on gold surface over either 3418A or amyl xanthate in the solution, rendering gold surface hydrophilic.

The present research investigates sodium hypochlorite, hydrogen peroxide, sodium metabisulfite and copper sulfate as activators to lessen the depressant effect of cyanide. The results indicate that the zinc recovery exceeded 93%, 90%, 85% and 95% at the dosages: sodium hypochlorite 1.5 ml/L, hydrogen peroxide 2 ml/L, sodium metabisulfite 1.67 × 10−3 mol/L and copper sulfate 2 × 10−3 mol/L, respectively. According to the results of FTIR spectrum and zeta potential, it is suggested that the studied marmatite was depressed due to the adsorption of CN− on the surfaces of marmatite by chemical bonds, and moreover the afore enhanced activators can present Zn C N bonds on the surfaces of depressed marmatite by oxidation or ionic adsorption, respectively, and hence activate the depressed marmatite for flotation.

In cyanidation process, the free cyanide concentration is the main variable that controls the efficiency of the process, and the plant operators determine it by silver nitrate titration. If sulfide ion is present in the solution the free cyanide concentration cannot be determined by titration as the silver added in the titration reacts with the sulfide ion to form a black solid (silver sulfide) that hinders the visual detection of the titration endpoint. This work proposes the potentiometric detection of the endpoint of the titration and demonstrates, thermodynamically and with analyses performed in synthetic solutions, that it is possible to determine both, sulfide and cyanide in the solution, even if the endpoint of the titration cannot be visually detected.

Thiosulphate, as an environmentally friendly lixiviant, is currently identified as the most promising alternative to cyanide for the extraction of gold from ores in sensitive locations. The Cu(II) ammine complexes acts as the catalytic oxidant to increase the thiosulphate leaching rate of gold, which also causes an undesirable high consumption of thiosulphate. Due to the complex solution chemistry of Cu–NH3–S2O32−–H2O system, fundamentally supported mechanisms to justify the catalytic effects and the consumption of thiosulphate are yet to be developed. Thermodynamic analysis offers essential information on which chemical species that could form and how these species influence the solution chemistry under various conditions. In this paper, predominance area and species distribution diagrams for gold leaching in the Cu–NH3–S2O32−–H2O system were constructed from a series of calculations using updated thermodynamic data. The diagrams reveal insights into the thermodynamics of gold dissolution, thiosulphate stability and redox behaviour of Cu(II)/Cu(I) couples. Analysis of the results presented in this paper can provide a theoretical basis for the complex thiosulphate leach solution chemistry that allows for a better understanding on the relationships between the redox behaviour of Cu(II)/Cu(I) couples and the behaviours of gold leaching and thiosulphate decomposition.

Gold cyanidation is an important hydrometallurgical route central to which are the gold/mineral galvanic interactions which pose perennial challenges to understanding gold dissolution. In this study, Density Functional Theory simulations have been performed to unveil some important electronic features of gold/pyrite associations on gold cyanidation. Three typical gold/pyrite occurrences have been considered, namely, model_1) gold slab/pyrite slab association to emulate contactless liberated gold and pyrite with up to a few A separations; model_2) gold-substituted pyrite solid-solution slab to emulate atomically dispersed gold in a sulfidic matrix; and model_3) pyrite slab with tethered gold nanocluster, in addition to isolated gold nanocluster and to pure pyrite slab. Gold/pyrite associations for all three studied models have been shown to induce electron transfer from gold to both pyrite’s sulfur and iron atoms. Model_1 revealed that galvanic interactions between liberated gold and pyrite can occur up to 10 A separation distance between the metal and the sulfide slabs. However, the efficacy of galvanic interactions decreased with increasing the inter-slab separation distances. Model_2 divulged that S-substitutive incorporation of Au in the pyrite structure resulted in expanded crystal dimensions as well as changes in electron properties of gold, iron and sulfur atoms in the solid-solution structure. Occurrence of gold nanocluster on pyrite surface (model_3) led to the formation of metal-metal Au-Fe bond and covalent Au-S bond. Sorption interactions of cyanide with gold and pyrite surfaces in presence of dioxygen were investigated with the three model variants. The DFT simulations showed that CN− preferentially adsorbed on gold instead of pyrite sites provided gold/pyrite galvanic interactions deprived gold from electron density at the expense of pyrite. The cyanide adsorption energies were found to follow the sequence: model_1 with substituted Au+ > model_2 > model_3 with substituted Au+ > model_1 (short distance) > model_3 > pure pyrite (1 0 0) & model_1 (long distance contact) > isolated Au9 nanocluster. As Au electron deficiency via galvanic interactions or gold oxidation was enhanced, gold-cyanide interaction energy increased portending the promotion of gold dissolution in cyanidation processes.

A synergistic leaching system using glycine containing starvation levels of cyanide as lixiviants has been shown to be an effective approach to leach gold-copper ores, allowing the consumption of cyanide to be reduced significantly while glycine is recycled. Sulfide precipitation to remove the bulk of the copper was studied. The previous study on the precipitation behaviour of Cu and Au from the alkaline glycine-cyanide solution shows that the cupric (Cu2+) glycinate can be easily precipitated, while the gold and cuprous (Cu+) cyanide species remain stable in the solution. Due to the sparingly soluble nature of metal sulfides, colloidal and poorly settling particles are usually formed without control methods, which create challenges for solid-liquid separation processes such as thickening and filtration. This study investigated the effects of chemical and operational conditions on particle characteristics particularly particle size distributions (PSD). Settling characteristics, particle morphologies and particle structure were also studied. In the presence of divalent cations such as Ca2+ and Mg2+, particularly Ca2+, large and fast settling particle agglomerates were generated. Increasing ionic strength of the solution was also noted to enlarge the particles. The high supersaturation level has insignificant effects on the PSD as long as Ca2+ is present. A relatively large particle size is generated at a medium stirring speed with fast addition rate. There are no significant effects of aging, heating, and seeding on the PSD, but these factors profoundly influenced the morphologies of the individual particles according to the SEM results. SEM and XRD analysis illustrate that a more mature and crystalline copper sulfide precipitates were produced after aging, heating, or seeding.