Two-step oxidation of a refractory gold-bearing sulfidic concentrate and the effect of organic nutrients on its biooxidation

Methods for improving the treatment efficiency of a refractory gold-bearing sulfidic concentrate are proposed. These methods consist of the oxidation of the concentrate during a two-step process, which includes a high temperature ferric leaching step and a subsequent biooxidation step, and the use of organic nutrients during the biooxidation step. The concentrate contained 34.7% pyrite and 7.9% arsenopyrite. The biooxidation of the concentrate (for a one-step process) was conducted at 45 °C in two bioreactors that were connected in series under continuous conditions. The pyrite and arsenopyrite oxidation levels after 240 h were 60.2% and 92.0%, and the gold recovery level by carbon-in-pulp cyanidation was 65.7%. The two-step process included the leaching of the concentrate by a biologically generated Fe³⁺-containing solution and the subsequent biooxidation of the leach residue. In this case, the pyrite and arsenopyrite oxidation levels after 240 h of biooxidation were 65.7% and 94.1%, and the gold recovery level was 71.7%.The effect of an organic nutrient (yeast extract) on biooxidation during the two-step process was studied. The pyrite and arsenopyrite oxidation levels after 240 h of biooxidation under mixotrophic conditions were 73.5% and 95.1%, and the gold recovery level was 77.9%. The effect of the organic nutrient on the microbial population was determined. Sulfobacillus thermosulfidooxidans and Acidithiobacillus caldus were the predominant microorganisms studied under both autotrophic and mixotrophic conditions. Archaeon Acidiplasma sp. MBA-1 was a minor component of the microbial community under autotrophic conditions but was one of the predominant microorganisms studied under mixotrophic conditions. These results suggest that the organic nutrient changed the composition and increased the activity of the microbial population.Thus, a two-step process with organic nutrients added during biooxidation may be considered as an effective strategy for treating refractory pyrite–arsenopyrite concentrates.     

Use of mesalime and electric arc furnace (EAF) dust as neutralising agents in biooxidation and their effects on gold recovery in subsequent cyanidation

The cost of lime/limestone for neutralisation is the second largest operating cost in bioleaching. Therefore, these studies have been conducted with the aim to investigate the possibilities for use of by-products such as mesalime and electric arc furnace (EAF) dust for neutralisation during biooxidation of a refractory gold concentrate. Experiments were carried out using a retention time of 57 h in a one-stage reactor and the influence of two industrial by-products on the biooxidation performance was evaluated. The neutralising capacity of EAF dust was lower, while the mesalime was similar to the Ca(OH)₂ reference. The arsenopyrite oxidation in experiments ranged from 85% to 90%, whereas the pyrite oxidation was 63–74%. In subsequent cyanidation, final gold recoveries of 90% were achieved in bioresidues from mesalime and Ca(OH)₂, while the EAF dust bioresidue had a recovery of 85%. A comparatively high elemental sulphur content in EAF dust probably encapsulates part of the gold, which explains the lower recovery for the EAF dust bioresidue despite a longer residence time. Cyanide consumption was relatively high and ranged from 8.1 to 9.2 kg/ton feed after 24 h of cyanidation. Overall, the by-products tested here have proved to be feasible options as neutralising agents in bioleaching operations.     

Release of arsenic from cyanidation tailings

At a gold mine in northern Sweden, gold occurring as inclusions in pyrrhotite and arsenopyrite is leached by cyanidation of the ore. The main sulphide minerals in the ore are pyrrhotite and arsenopyrite. Effluents from the cyanidation process are treated with Fe₂(SO₄)₃ to form Fe-precipitates suitable for the co-precipitation of As. The aim of this study was to perform static and kinetic leaching tests on the ore and tailings to define geochemical processes governing As mobility. Sequential leaching tests suggested that the majority of dissolved As deriving from the sulphide fraction in the ore was incorporated in newly formed Fe-precipitates in the tailings. The mobility of As in the tailings was therefore mainly dependent on the stability of these As-bearing Fe-precipitates. Weathering cell tests (WCT) involving 31 weekly cycles of wetting and air exposure were conducted to assess the stability of the As in the tailings under accelerated weathering conditions. The first stage of the WCT was characterized by a pH ≈ 5 and low As leaching, probably driven by the dissolution of amorphous Fe-As species. In the second stage of the WCT, leaching of Fe, S and As increased and the pH decreased to <3.5. An increase of the leachate’s molar Fe/S-ratio suggested that pyrrhotite oxidation was occurring. The falling pH destabilized As-bearing Fe-precipitates, causing further As release. The total As release during the WCT corresponded to only a small proportion of the tailings’ total As content. The accelerated As-leaching observed towards the end of the WCT could thus indicate that its release could increase progressively over time.     

Leaching and CIL processes in gold recovery from refractory ore with thiourea solutions

In this study, the applicability of leaching and CIL processes in gold recovery with thiourea method, alternative to the cyanidation from the refractory Gümüşhane-Kaletaş/Eastern Black Sea Region (Turkey) ore was investigated.The experiments were conducted at laboratory conditions using ore samples of which approximately 80% were ground to −0.038 mm. The grade of the ore samples was 6.8 g Au/ton. At the first part of the experimental studies, assuming that the gold could be recovered with CIC and CIP processes, the effects of pH, thiourea, oxidizing agent consumption, and leaching time on leaching were investigated. Then, on the basis of the optimum pH and reagent consumption values obtained in the first part (pH = 1.5, 15.2 kg thiourea/ton ore, 140.9 kg iron(III) sulfate/ton ore and 46.2 kg sulfuric acid consumption/ton ore) and adding 50 kg activated carbon/ton ore at the beginning of experiments, the gold leaching extents were obtained for the same leaching times. At this part, the applicability of CIL process in gold recovery with thiourea was investigated for the first time. As a result of the experiments, although higher gold leaching extents were obtained in CIL process, the increase in extent was about maximum 8% and the highest gold leaching extent was obtained as 75% at the end of the 5th hour.     

The influence of pyrite pre-oxidation on gold recovery by cyanidation

The influence of pyrite pre-oxidation in alkaline solutions on gold recovery by cyanidation from Twin Creek refractory gold ore in which pyrite was identified as the major sulfide mineral has been investigated with the aid of electrochemical measurements, leaching experiments, and direct analysis of reaction products for selected residues. It was found that gold recovery by cyanidation in bottle roll experiments mainly depended on the extent of pyrite pre-oxidation. The rate of pyrite oxidation in alkaline solutions measured by electrochemical measurements, including chronoamperometry and linear sweep voltammetry, increased with an increase in pH, potential, and temperature. All alkaline reagents used for the electrochemical measurements, NaOH, NH₄OH, Na₂CO₃ and Ca(OH)₂, showed a similar effect on pyrite oxidation kinetics. However, the results of alkaline pre-oxidation for pyrite of the Twin Creek refractory gold ore suggested that NaOH and Na₂CO₃/Ca(OH)₂ were superior to Ca(OH)₂. Without pre-oxidation, cyanide leachable gold was found to be only 20% which could be increased to 70% under appropriate pre-oxidation conditions. At the same time, cyanide consumption decreased from 2.5 kg/t ore to 1.5 kg/t ore.Selected residues after pre-oxidation and cyanidation were examined by X-ray diffraction. Backscattered electron images of pyrite particles in these residues were taken. The reaction products at the surface of pyrite particles were found to be iron-, silicon-, and calcium-bearing compounds with variable amounts of sulfur as determined by X-ray energy dispersion analysis. Additionally, some mineral fines, such as aluminum and/or potassium-bearing minerals, were found to be present at the partially oxidized pyrite surface.     

Dissolution of gold during pyrite oxidation reaction

Numerous papers discuss the mechanism of alkaline oxidation of pyrite but there is limited information available describing the actual kinetics of the pyrite sulphide to thiosulphate reaction. A previous investigation in this series determined the rate of sulphide sulphur oxidation and thiosulphate yield in the reaction of pyrite with sodium hydroxide under various testing conditions. The goal of the current study is to validate these rates using two different gold-containing pyrite concentrates. A further objective of the current work is to investigate the simultaneous dissolution of gold with in situ formed thiosulphate during pyrite oxidation.It was found that at 20 psi oxygen overpressure and a temperature of 80 °C, the initial rate of sulphide oxidation and thiosulphate yield were close to 0.08 mol/h and 0.0155 mol/h, respectively. These rates are in agreement with previously published data. However, a shift from linearity occurred when the pH decreased below 12. A rapid decay of thiosulphate was evidenced at pH 8.3–9.2 while E_H was in the range of 22–141 mV. Based on relevant thermodynamic analysis of metastable thiosalts system, such rapid decomposition is not expected at these pH and E_H values. It is believed that the presence of unreacted pyrite acting as a catalyst caused this behaviour. It appears that under mildly alkaline conditions, the rate of oxidation of sulphide to thiosulphate becomes slower than the rate of thiosulphate degradation, which causes a net loss of thiosulphate in the system. The maximum extraction of gold and silver (96% and 75% respectively) was achieved under conditions of pH < 12.     

A study of atmospheric acid leaching of a South African nickel laterite

Nickel and cobalt acid leaching from a low-grade South African saprolitic laterite using sulphuric acid was studied. Ore characterisation was performed by XRD and XRF. Batch agitation leaching tests were conducted at atmospheric pressure investigating main parameters: particle size and percent solids at 25 °C and 90 °C. Ore characterisation showed that the ore is a saprolitic laterite with nickel present in lizardite. Leaching tests showed that nickel and cobalt could be leached from the ore at atmospheric pressure. Nickel was found to be more leachable from the coarser −106 + 75 μm fraction, with 98% Ni being extracted at 90 °C after 480 min. Cobalt was not favoured by variation in particle size and increased percent solids. Increasing ore percent solids improved nickel extraction at 25 °C however at 90 °C extraction decreased due to a diffusion layer build-up as a result of amorphous colloidal silica. The co-dissolution of magnesium and iron was elucidated. Nickel leaching data at increased temperature and percent solids fit the shrinking core model equation, k_dt = 1−2/3x − (1 − x)^2/3 showing that nickel leaching reaction was diffusion controlled under the set conditions.     

Combined bio-hydrometallurgical process for gold recovery from refractory stibnite

The aim of this work concerns the recovery of gold by a combined biometallurgical and hydrometallurgical treatment. The biooxidation of a refractory gold-bearing stibnite (antimony sulphide) favours the increase of gold recovery during the subsequent hydrometallurgical process (cyanidation, adsorption onto activated carbon, desorption from activated carbon and electrowinning).Tests were conducted at laboratory scale utilising a refractory stibnite ore. The gold content of the sample was 32 g/t Au. Bacterial cultures utilised in the biological test consisted predominantly of Thiobacillus genus. After a long leach time (24 hours) low recovery was obtained by direct cyanidation (less than 4% Au), while a gold extraction yield of about 80% was achieved combining biometallurgical and hydrometallurgical process.Experimental results have shown the technical feasibility of the biometallurgical pretreatment prior to the hydrometallurgical process. A complete process flow-sheet has been outlined.     

Thiosulphate leaching of gold in the presence of ethylenediaminetetraacetic acid (EDTA)

The addition of low levels of ethylenediaminetetraacetic acid (EDTA) in the ammoniacal thiosulphate gold leach system lowered the catalytic cupric/cuprous redox equilibrium potential, hence the mixed solution potential and reduced the consumption of thiosulphate. In the leaching of pure gold, gold dissolution was enhanced in the presence of EDTA at a relatively low concentration, but excessive EDTA decreased gold dissolution. Raman analysis of the leached gold foil indicated that the stabilisation of thiosulphate by EDTA decreased the formation of the passivation layers of elemental sulphur and copper sulphide at the gold surface. In the leaching of a sulphide ore, the leaching kinetics and overall extractions of gold and silver were enhanced substantially, while the consumption of ammonium thiosulphate was reduced from 9.63 kg/t to 3.85 kg/t in the presence of 2.0 mM EDTA after 24 h leaching. This beneficial effect became more pronounced at a higher EDTA concentration. The enhanced gold and silver extractions by EDTA were attributed to the increase in the dissolution of gold and silver bearing sulphides, the stabilisation of copper and thiosulphate in leach solutions, the prevention of leaching passivation and the decrease in the interference of foreign heavy metal ions. The use of EDTA in the ammoniacal thiosulphate leaching system makes it practical to achieve satisfactory gold extraction over extended periods of leaching under low reagent concentrations, where the consumption of thiosulphate is low.     

Selective sulphidation and flotation of nickel from a nickeliferous laterite ore

The sulphidation of a nickeliferous lateritic ore was studied at temperatures between 450 and 1100 °C and for sulphur additions of 25–1000 kg of sulphur per tonne of ore. The experiments demonstrated that the nickel could be selectively sulphidized to form a nickel–iron sulphide. It was found that both the grade and the sulphidation degree largely depended upon the temperature and the sulphur additions, with temperatures above 550 °C exhibiting the highest nickel sulphidation degrees and grades. A DTA/TGA with mass spectrometer was used to further elucidate the nature of the phase transformations that occurred upon heating of the ore in the presence of sulphur.It was found that at low temperatures, the Fe–Ni–S phase was submicron in nature and heating to temperatures between 1050 and 1100 °C allowed for the growth of the particles, due to the increased sulphide mobility associated with the formation of a liquid sulphide matte phase, containing dissolved oxygen. Flotation studies conducted on 60 g samples showed that the sulphides responded to flotation with maximum grades of up to 6–7 wt.% nickel being achieved. Recoveries were approximately 50% on a sulphide basis and it was determined that the low nickel grades were due to the entrainment of magnetite fines.     

Evaluation of leaching parameters for a refractory gold ore containing aurostibite and antimony minerals: Part I – Central zone

A cyanidation study was conducted on a mild refractory gold ore sample from the Central zone of Clarence Stream Property, owned by Freewest Resources Canada, to develop a leaching strategy to extract gold. Gold, at a grade of 8.00 g/t, is present as native gold, electrum and aurostibite. The ore also contains 2.8% pyrrhotite, together with several antimony minerals (0.8% berthierite and gudmundite, 0.18% native antimony and stibnite). It also exhibits weak preg-robbing properties with 0.16% organic carbon. Aurostibite, a gold antimony compound, is particularly known to be insoluble in cyanide solution. The antimony dissolves in cyanide solution to form antimonates, which retards gold dissolution. Industrial practice of extracting gold from aurostibite generally consists of producing a flotation concentrate, which is leached in a pipe reactor at low alkalinity and high oxygen pressure with about 20 g/L cyanide.The proposed new approach is efficient and allows the extraction of gold directly from an ore at atmospheric pressure and a low cyanide concentration at pH 10.5. The effects of grinding, pre-treatment, lead nitrate, kerosene and cyanide concentrations have been investigated. The maximum gold extraction obtained on the ore was 87.9% using 800 ppm NaCN, 500 g/t lead nitrate, 30 g/t kerosene, DO (dissolved oxygen) 10 ppm and pH 10.5 in 168 h. The associated cyanide consumption was 1.3 kg/t. The additions of lead nitrate and kerosene increased gold extraction. In comparison to a P₈₀ of 74 μm, a P₈₀ of 30 μm significantly increased gold extraction. Gold in solid solution in gudmundite and arsenopyrite was believed to be responsible for the un-leached fraction until mineralogical analysis of hydroseparation concentrates of leach residues showed that most of the un-leached gold occurs as aurostibite, either as locked grains in sulphides/sulpharsenides or as grains with passivation rims of an Au–Sb–O phase. Coarse gold was also found. Gold extraction was not sensitive to cyanide concentration from 250 to 1200 ppm NaCN and high pH was detrimental. Decreasing the cyanide concentration reduced the cyanide consumption from 1.39 to 0.85 kg/t. The removal of coarse gold using a Knelson concentrator and a Mosley table prior to leaching increased the gold extraction to 90.4% (leach residue at 0.77 g/t).     

Mineral sulfide concentrate leaching in high temperature bioreactors

Mineral sulfide concentrate oxidation by thermoacidophilic archaea was reported in the 1970s with a culture which was described as “possibly related to Sulfolobus” but which contained the species later named as Acidianus brierleyi. In the following two decades, the mineral sulfide-oxidizing capacity of Sulfolobus metallicus became the subject of many studies partly because it appeared to tolerate higher concentrations of copper and solids in bioreactors than those tolerated by the strains of Acidianus and Metallosphaera that were available. However, some bioreactor pilot plant developments followed with mixed cultures dominated by A. brierleyi at 70 °C. At higher temperatures, pilot and industrial demonstration plants used mixed cultures dominated by poorly characterized novel species. Most of these novel species will require classification in novel genera to reflect their phylogenetic separation from previously named thermoacidophilic archaea.     

Characterisation of coarse composite sphalerite particles with respect to flotation

The flotation response of a typical zinc-lead (Zn/Pb) ore, with respect to coarse composite (sulphide/non-sulphide) particles is reported. The flotation tests were carried out on a selected feed particle size range (−600 + 75 μm, at P₈₀ of 390 μm) and the recovery of Zn composite particles analysed on a size by size basis. The best results were achieved with the use of 75 g/t sodium isopropyl xanthate (SIPX), obtaining a Zn recovery of 77%, with a significant improvement at the coarse end of the particle size distribution. Computerised scanning electron microscope (QEMSCAN) was used to characterise value mineral grain size and degree of liberation, as well as gangue and sphalerite association in particles reporting to both concentrate and tailings. A new characterisation function (Locking ratio, LR) was developed based on the data from the automated mineralogical analysis to characterise particles into two-phase composites with different degree of locking texture (simple and complex). The function, which is based on the mode of occurrence of sphalerite, grain size, proportion and composition of the constituent minerals in each particle, was used to study the flotation response of the particles with different degrees of locking. The results highlight the difference in recoverability of the sphalerite bearing particles with different degrees of locking, with simple locking texture giving higher recovery than complex locking texture, for the same overall liberation.     

Copper–gold ore processing with ion exchange and SART technology

Anglo Asian Mining has developed a 50,000 oz Au/yr open pit gold mine at Gedabek in Western Azerbaijan. The deposit at Gedabek is a copper–gold porphyry, comprising both oxide and sulphide ore mineralisation, which is being mined at the rate of about 1 million tons of ore per year. Ore processing is by conventional cyanide heap leaching, which produces a pregnant leach solution (PLS) containing 1–2 ppm of gold, together with 1000 ppm or more of copper. The PLS is treated by column ion exchange, using Dow’s gold-selective MINIX resin. Loaded resin is stripped with an acidic thiourea solution, from which gold and silver are electrowon on to stainless steel mesh cathodes. Copper concentrations in the leach solutions are controlled by passing part of the PLS flow through a SART process, where the acronym stands for “Sulphidisation, Acidification, Recycling and Thickening”. The product from the SART process is a copper/silver sulphide precipitate, which is thickened, filtered and dried and then sold for copper smelting.     

A screen of some native Australian flora and exotic agricultural species for their potential application in cyanide-induced phytoextraction of gold

The phytoextraction of gold is an exciting new area of research involving the use of plants to extract gold from a low grade ore and waste products. The aim of this study was to screen some Australian native plant species and exotic agricultural species for their potential use in cyanide-induced phytoextraction of gold from a crushed ore body. Plants were grown in an ore body with a gold grade of 1.75 g/ton gold and treated with 0.1 g/kg and 1 g/kg cyanide as aqueous NaCN. The above-ground biomass was harvested one week after treatment and the gold concentration determined. Mean gold concentrations of up to 27 g/ton (d.w.) were found in the plant tissue of Trifolium repens cv. Prestige. Higher gold concentrations were generally found in the stems and older leaves with the lowest gold concentrations in the young leaves. Plants grown in the 1 g/kg cyanide treatment generally had higher concentrations of gold in the plant tissue than plants grown in the 0.1 g/kg cyanide treatment. This study demonstrates the potential use of plants to extract and concentrate gold from a low grade ore and waste products.     

Evaluating waste rock mineralogy and microtexture during kinetic testing for improved acid rock drainage prediction

This study integrates detailed mineralogical and microtextural analyses of waste rock with the results of standard kinetic test procedures to identify the mineralogical changes that influence leachate chemistry over time. The integration of mineralogy and texture provides the opportunity for improved mine waste management strategies and acid rock drainage (ARD) prediction.Waste rock material from an abandoned gold mine in northern Queensland, Australia, was subjected to column leach kinetic testing over a 30 week period. The column feed comprised of a range of waste rock lithologies (porphyritic rhyolite, massive arsenopyrite, massive pyrite ± galena, and semi-massive polysulphide). In total, 12 individual columns were established to represent six lithologies prepared to two different size fractions (−10 mm and −4 mm). The mineralogy and microtextural characteristics of the column feed material was defined using quantitative X-ray diffractometry (QXRD), scanning electron microscopy and laser ablation (LA-ICPMS) at the start of kinetic tests, and at 5 week intervals during the length of the tests. These data were directly correlated with leachate chemistry (i.e., pH, SO₄ and select elements).Results of this study indicated that sulphide oxidation was strongly influenced by the morphology of sulphide minerals, their trace element contents, the presence of mineral micro-inclusions and galvanic interactions with other sulphide minerals. Waste rock with abundant arsenopyrite was consistently the most acid forming, and oxidised to scorodite (enriched in Zn, Pb and Cu). Pyrite was commonly As-rich as indicated by LA-ICPMS mapping. QXRD results indicated that the abundance of rhomboclase, jarosite, alunite and hydrous ferric oxides increased over time. Galena weathered rapidly to porous anglesite, particularly when in direct physical contact with pyrite. Sphalerite contents decreased consistently over the 30 weeks implying its oxidation, however few reaction products were directly observed. By week 30, the −4 mm fraction material generated lower pH leachate, higher mass release of elements and sulphate for the majority of samples. This indicates that the particle size used in kinetic tests can exert a significant control on leachate chemistry, especially in the absence of abundant neutralising minerals. This contribution demonstrates the value of integrating mineralogy and microtextural analyses during kinetic testwork to improve the interpretation of sulphide oxidation for better prediction of ARD.     

Selective elution of the gold cyanide complex from anion exchange resin using mixed solvents

This study is concerned with the use of mixed solvents for the elution of the cyanide complexes of copper and gold from Purolite A500, a strong-base anion exchange resin. The mixed solvents investigated include acetone + water, dimethylsulfoxide + water and N-methyl-2-pyrrolidone + water. Three types of counterions are employed in each of the mixed solvents: CN⁻, Cl⁻ and OH⁻. The effects of counterion concentration and mixed solvent composition on the elution of the complexes are examined. High recoveries of the gold cyanide complex are achieved in the mixed solvents at relatively low counterion concentrations. In contrast, the recoveries of the copper cyanide complexes are 1–3 orders of magnitude lower for the given initial loading of the metals on the resin. The selectivity of the elution process for gold is discussed in terms of the degree of solvation of the various anions in the mixed solvents. The results of this study point to the possibility of using mixed solvents to develop an elution process that is selective for gold over multivalent cyanide complexes.     

Recovery of indium from sphalerite ore and flotation tailings by bioleaching and subsequent precipitation processes

Flotation tailings dump material of the former lead–zinc mine near Freiberg (Germany) consists of fine grained quartz, feldspar, mica as well as the sulphide minerals pyrite, galena and sphalerite not recovered by flotation. Sphalerite contains, aside from iron, copper and cadmium, significant amounts of indium (up to 0.38% (w/w)) leading to indium contents up to 70 mg/kg in the mine tailings. Preliminary thermodynamic assessment showed a comparatively small Eh–pH-range where bioleaching is possible and indium is not hydrolytically precipitated. Shake flask bioleaching of original polymetallic sphalerite ore from the Freiberg mining district (400 mg/kg indium) showed maximum zinc and indium recovery rates of almost 100% or 80%, respectively. First bioleaching tests on tailings material achieved zinc and indium yields of up to 80%. A stepwise precipitation process is being developed for indium recovery from the PLS (pregnant leaching solution) consisting of combined iron/indium precipitation and subsequent processing of the indium pre-precipitate.     

A Pilot Scale Study of Cenosphere Recovery and Concentration using the Inverted Reflux Classifier

Cenospheres are hollow spherical particles formed as part of the fly ash waste of coal-fired power stations. In a previous paper Kiani et al. (2015) investigated the recovery and the concentration of these particles using an Inverted Reflux Classifier (IRC) at a laboratory scale, of cross-section 0.100 m × 0.086 m, achieving a throughput advantage over a conventional fluidized bed by a factor of 54. The present paper investigated the potential to achieve scale-up, utilizing a pilot scale device with cross-section 0.3 m × 0.3 m. The product grade and recovery were examined as a function of the solids yield by varying the product volumetric rate relative to the feed volumetric rate. The performance data were compared directly with those obtained at the smaller laboratory scale. Agreement was excellent. The performance was also examined as a function of the feed slurry flux, with good agreement again evident at the laboratory and pilot scales. Overall, the separation performance was excellent, with a cenosphere recovery of about 80% achievable at a high upgrade of 19 while a recovery of 75% was achieved at an upgrade of 38. Here the feed solids flux was 4.2 t/(m² h). It is noted that much higher upgrade was achieved at a recovery of about 80% in the former study by operating at a lower solids feed flux. This paper provides the necessary basis for proceeding with a full scale implementation of this technology.     

Simultaneous sulfide oxidation and gold leaching of a refractory gold concentrate by chloride–hypochlorite solution

In this research, oxidation of sulfide and leaching of gold from a gold–bearing sulfide concentrate using chloride–hypochlorite solution has been investigated. Effects of calcium hypochlorite concentration, initial pH and sodium chloride concentration on the recovery of gold were examined. Two conditions were considered; the stability range of the gold complex (Eh > 900 mV) and formation of chlorine gas (pH < 3.5). During leaching, due to oxidation of sulfide and generation of acid, pH dropped. About 82% of gold was extracted from 200 g/L concentrate after 2 h using 200 g/L Ca(OCl)₂, 200 g/L NaCl at initial pH of 11, stirring speed of 600 rpm and temperature of 25 °C.