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.     

Copper leaching from chalcopyrite concentrate in Cu(II)/Fe(III) chloride system

This study was conducted to develop a novel process for copper recovery from chalcopyrite by chloride leaching, simultaneous cuprous oxidation and cupric solvent extraction to transfer copper to a conventional sulfate electrowinning circuit, and hematite precipitation to reject iron. Copper leaching from chalcopyrite concentrate in ferric and cupric chloride system was investigated using a two-stage countercurrent leach circuit under a nitrogen atmosphere at 97 °C to minimize the concentrations of cupric and ferric ions in pregnant leach solution for subsequent copper solvent extraction while maintaining a maximum copper extraction. A high calcium chloride concentration (110–165 g/L) was used to maintain a high cuprous solubility and enhance copper leaching. With 3–4 h of leaching time for each stage, the copper extraction reached 99% or higher while that of iron was around 90%. With decreasing concentrate particle size from p80 of 26 to 15 μm, the copper extraction increased by about 0.2% while the iron extraction increased by about 2.0%. The concentration of Cu(II) + Fe(III) in the pregnant leach solution was able to be reduced to 0.04 M. When the cupric concentration fell below the above limiting value, the elemental sulfur present was reduced by cuprous ions to form copper sulfide, eventually stopping the leaching of copper. Under this condition, only iron was leached. A very small amount of sulfur (1.2–1.4%) was oxidized to sulfate, resulting in an increase from 3 to 9 g/L in HCl concentration. The extractions of trace metals (Cr, Pb, Ni, Ag and Zn) were 96–100%.     

Extraction of indium from zinc plant residues

The main purpose of this study was to extract indium from the Irankoh zinc plant residue. The Irankoh zinc plant residue contained 145 ppm indium. The optimum conditions for leaching of indium and reduction of ferric ion in reductive leaching were obtained at temperature of 90 °C for a leaching duration of 3 h with sulfuric acid concentration of 100 g/L and the amount of required sodium sulfide for reduction of ferric was 1.5 times of stoichiometric quantity of iron. Then, to prepare concentrated indium solution, indium was selectively precipitated from the leach solution. The pH of leach solution was adjusted to 6 with ammonia solution in 90 °C for selective indium precipitation, and reaction time was considered to be 10 min. Then the resulting precipitation was dissolved using hot sulfuric acid solution, and the solution was subject to solvent extraction and cementation using zinc powder to recover indium.     

Co-processing of sulfidic mining wastes and metal-rich post-consumer wastes by biohydrometallurgy

The consequence of a strong economic growth in emerging countries combined with the rise of the world population is an increase in the demand for raw materials, leading to growing concern regarding their availability and the global efficiency of the supply chain. These tensions reinforce the need to associate the development of the recycling industry to the identification of new resources which could be used for the recovery of valuable materials. The purpose of this study is to develop a novel biological co-processing approach for the recovery of strategic metals in both sulfidic mining wastes and post-consumer wastes (WEEE). The principle of this treatment is based on two steps: mine wastes are biologically oxidized, resulting in the production of a ferric iron-sulfuric acid lixiviant solution which is used to leach base and other soluble metals contained in e-scraps. Batch tests were carried out using flotation tailings wastes containing 60% of pyrite and grinded Printed Circuit Boards (PCB < 750 μm) with a solid load of 2.5%. Two series of tests were conducted in order to study the influence of the ferric iron concentration and of the bacterial activity on metals dissolution. Results showed that a higher ferric iron concentration led to an increase in the dissolution rate of copper which is the main metal contained in the PCBs. Moreover, a dissolution yield of 98.3% was reached for copper after 2 days when bacterial activity was observed, corresponding to an increase of about 20% compared to the tests without bacterial activity. Finally, this study highlights the importance of the availability of ferric iron and of the bacterial oxidation of ferrous iron for the feasibility of this bioleaching process dealing with the recycling of PCBs.     

Simulated small-scale pilot plant heap leaching of low-grade oxide zinc ore with integrated selective extraction of zinc

The leaching of low-grade oxide zinc ore and simultaneous integrated selective extraction of zinc were investigated using a small-scale leaching column and laboratory scale box mixer-settlers. Di-2-ethylhexyl phosphoric acid (D2EHPA) dissolved in kerosene was used as an extractant. The results showed that it was possible to selectively leach zinc from the ores by heap leaching. The zinc concentration of the leach liquor in the first leaching–extraction circuit was 32.57 g/L, and in the 16th cycle the zinc concentration was 8.27 g/L after the solvent extraction. The leach liquor was subjected to solvent extraction, scrubbing and selective stripping for the enrichment of zinc and the removal of impurities. The pregnant zinc sulfate solution produced from the stripping cycle was suitable for zinc electrowinning.     

Recovery of zinc and cadmium from industrial waste by leaching/cementation

Metallic zinc production from sulfide zinc ore is comprised by the stages of ore concentration, roasting, leaching, liquor purification, electrolysis and melting. During the leaching stage with sulfuric acid, other metals present in the ore in addition to zinc are also leached. The sulfuric liquor obtained in the leaching step is purified through impurities cementation. This step produces a residue with a high content of zinc, cadmium and copper, in addition to lead, cobalt and nickel. This paper describes the study of selective dissolution of zinc and cadmium present in the residue, followed by the segregation of those metals by cementation. The actual sulfuric solution, depleted from the electrolysis stage of metallic zinc production, was used as leaching agent. Once the leaching process variables were optimized, a liquor containing 141 g/L Zn, 53 g/L Cd, 0.002 g/L Cu, 0.01 g/L Co and 0.003 g/L Ni was obtained from a residue containing 30 wt.% Zn, 26 wt.% Cd, 7 wt.% Cu, 0.35 wt.% Co and 0.32 wt.% Ni. The residue mass reduction exceeded 80 wt.%. Cementation studies investigated the influence of temperature, reaction time, zinc concentration in feeding solution, pH of feeding solution and metallic zinc excess. After that such variables were optimized, more than 99.9% of cadmium present in liquor was recovered in the form of metallic cadmium with 97 wt.% purity. A filtrate (ZnSO₄ solution) containing 150 g/L Zn and 0.005 g/L Cd capable of feeding the electrolysis zinc stage was also obtained.     

Investigating heap bioleaching: Effect of feed iron concentration on bioleaching performance

This paper describes an investigation into the effect of iron concentration in the leach solution on the bioleaching of a low grade copper ore, where chalcopyrite was the dominant copper sulphide. The concentration of dissolved iron is primarily controlled by pH and the relative proportion of ferric to ferrous iron, with significant jarosite precipitation occurring above pH ≈ 1.8 in a highly oxidised system. The solution pH may be increased by the dissolution of acid soluble gangue and when iron oxidation is significantly higher than sulphur oxidation. The study was approached using two experimental systems. In the former, the leach solution was recycled through an ore bed of low aspect (reactor height divided by diameter) ratio for a portion of the experiment. During the recycle phase, no acid was added to the system and acid consumption by gangue material led to a pH increase (1.6–2.2). The resulting jarosite precipitation reduced soluble iron from 2.5 g/l to less than 250 mg/l. Copper recovery decreased, but not in proportion to the decrease in iron. This was partly attributed to adsorption on, or entrainment within, the jarosites. To study the effect of reduced iron concentration on leach performance under more controlled conditions, bioleaching was performed in packed bed column reactors with feed iron concentrations ranging from 5 g/l to 200 mg/l. Observations indicated an initial decreased rate of copper liberation with reduced iron concentration in the feed. The relationship between available Fe³⁺ concentration and copper liberation was not proportional. However, with time, the liberation of copper became independent of iron concentration in the percolation liquor. Further, the specific rate of copper liberation was consistently below the theoretical value on a basis of ferric iron concentration. The highest values of copper liberation were reported at the lowest iron concentrations. In summary, while increased iron concentration in solution may enhance the initial rate of leaching, mineral availability appears to dominate CuFeS₂ leach kinetics through the majority of the leach. Furthermore, high iron concentrations in solution aggravate jarosite formation with concomitant retention of copper in the ore bed.     

Ferric leaching of copper slag flotation tailings

The pyrometallurgical production of copper generates slags, a residue with a significant content of this metal. Copper can be recovered from the slags by froth flotation after cooling, crushing, and grinding. The obtained Cu-concentrate is sent to the pyrometallurgical process. If grinding is not fine enough for efficient flotation, copper is lost in tailings. In this paper, the ferric leaching of slag flotation tailings is studied. Copper extractions of 66% are achieved by ferric leaching, and Cu content in tailings is reduced from 0.78% to 0.24%.     

The effect of redox control on the continuous bioleaching of chalcopyrite concentrate

A continuous bioleaching process was developed for the dissolution of chalcopyrite concentrate with electrochemically redox control. Therefore, using a flotation concentrate containing 46% chalcopyrite and 23% pyrite, bioleaching tests were carried out at 47 °C with 15% pulp density under controlled and uncontrolled redox conditions. To increase the copper recovery in contrast to the conventional bioleaching (∼39.62%), the effect of redox potential on the chalcopyrite bioleaching was investigated by electrochemically controlled bioleaching. The results showed that by controlling the redox potential, faster copper leach kinetics could be achieved. At last, reducing the redox potential from high levels to optimum window (420–440 mV SCE) caused an increase in copper recovery from around 39% to higher than 69% (over 25 g/L Cu²⁺).     

Evaluation of the response of valuable and gangue minerals on a recovery,size and liberation basis for a low-grade silver ore

Mineralogical characterisation is typically used to assist in the development of processing strategies for ores. This paper describes the application of mineralogical characterisation techniques in the development of flotation strategies for processing an ore from a low-grade silver deposit that contains a variety of rock types that have undergone hydrothermal alteration where zinc, lead and silver sulphides are typically the primary minerals of economic interest. Comprehensive mineralogical characterisation of the feed, concentrates and tailings from batch flotation tests was undertaken using both the mineral liberation analyser (MLA) and laser ablation inductively-coupled plasma mass spectroscopy (LA-ICP-MS). The results of mineralogical characterisation of the feed (head grade, 116 ppm) indicated that the majority of the silver occurred as solid solution in pyrite which assisted in the development of the flotation strategy used for this ore which resulted in approximately 87% of the total silver being recovered to rougher concentrate at a grade of 485 ppm.     

Ferric sulphate/chloride leaching of zinc and minor elements from a sphalerite concentrate

Atmospheric leaching of a sphalerite concentrate in sulphate and chloride media was performed and the effect of several variables, such as solid/liquid ratio and oxidant (Fe(III)) concentration were investigated. The behaviour of minor elements, such as Cu, In, As, Sb, Bi, Sn and Pb, was also studied under different conditions. The results showed that using a solid/liquid ratio of 5% (w/v) it was possible to leach 95% of zinc after 2 h, with a solution of 0.5 M H₂SO₄ and Fe₂(SO₄)₃ at 80 °C. The minor elements As, Sb and Bi were also completely leached whereas copper leaching was favoured by the use of chloride medium. The oxidation of Fe(II) during the leaching tests was studied and an improvement of 20% zinc extraction was observed in an oxygenated system. Cross-current leach tests using two/three stages and a solid/liquid ratio of 10% (w/v) were performed to achieve 90% of zinc extraction. The electron microprobe analysis of the leaching residues showed no change on the sphalerite composition after the leaching, which indicates that the leaching of sphalerite involves the break down of the sulphide structure.     

Development of a process for the separation of zinc and copper from sulfuric liquor obtained from the leaching of an industrial residue by solvent extraction

This paper describes a study of the separation of zinc and copper from the leach liquor generated in the treatment of the zinc residue (29.6 g/L Zn and 37.4 g/L Cu) by liquid–liquid extraction. In it, the influence of the extractant type and concentration, aqueous phase acidity, contact time and stripping agent concentration were investigated. Organophosphorus extractants (D2EHPA, IONQUEST®801 and CYANEX®272) and the chelating extractants (LIX®63, LIX®984N and LIX®612N-LV) were also investigated. The organophosphorus reagents are selective for zinc, while the chelating extractants are selective for copper. In the experiment, D2EHPA was found to be the best extractant. A sulfuric acid solution was used in the stripping study. Five continuous experiments were carried out until an optimal condition for the separation of the metals Zn and Cu was achieved. Experiment 5 was carried out in three extraction steps, three scrubbing stages and five stripping stages. In this experiment, a pregnant strip solution containing 125 g/L Zn and 0.01 g/L Cu was obtained and the concentration of the metals in the raffinate was 28.3 g/L Cu and 0.49 g/L Zn.     

Textural,mineralogical and chemical characteristics of copper reverb furnace smelter slag of the Okiep Copper District,South Africa

The Okiep Copper District in South Africa has produced more than 110 million tons at a grade of 1.71% Cu from several small mafic ore bodies. The ore was smelted on site and generated ∼5 mt of slag. During the life of mine attempts to recover copper from the slag by flotation had limited success. After mine closure the challenge of environmental rehabilitation and the possible disposal of the slag, triggered a reinvestigation into the viability of slag as a copper resource. Characterisation of the slag as a contribution to the potential copper recovery is the objective of this study.The slags are hard, vitreous with a matrix of Si–Fe–Al–Mg–Ca glass and laths of Mg–Fe–olivine, Fe–Mg–orthopyroxene and minor Cr-spinel. Copper grade varies between 0.11% and 0.42% with minor nickel, cobalt, molybdenum, zinc and tungsten. All economic elements are hosted by disseminated spheroidal prills which consist mainly of the copper sulphides bornite, chalcocite, covellite and chalcopyrite with exsolved sulphide phases of the minor base metals as well as rhenium and silver. Prills consisting of metallic copper and alloys are minor constituents. Prill diameter is highly variable with most in the 40–60 μm range and the historically poor copper recovery is attributed to the small prill size. Crushing of slag to −45 μm as opposed to the previous −75 μm should significantly increase sulphide liberation and recovery of copper and minor base metal sulphides by conventional flotation.Provided the operation is economically viable, redistribution of the processed slag to environmentally acceptable sites will resolve the present pollution and rehabilitation challenge related to the dumps in the Okiep Copper District. The operation will also have a positive socio-economic impact on this poverty-stricken part of South Africa.     

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.     

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.     

Recovery of dissolved platinum group metals from copper sulphate leach solutions by precipitation

The pregnant leach solution produced in the final leaching stage of base metal refineries (BMRs) operated by platinum producers contains impurities such as selenium and tellurium as well as other precious metals (OPMs, which include Rh, Ru and Ir). The aim of this project was to propose operating conditions for a thio-urea precipitation process that would allow maximum OPM recovery and impurity precipitation from the leach solution with minimal copper and nickel co-precipitation. Experimental results illustrating the effects that operating temperature (80 °C and 160 °C), pressure (atmospheric pressure and seven bar), stirring rate (250 rpm and 500 rpm) and thio-urea quantity (200% and 320% excess) have on the precipitation behaviour are presented.Virtually all of the Rh contained in the solution was precipitated irrespective of the values of the process variables studied. The maximum percentage Ru and Ir precipitation achieved were 87% and 60%, respectively. Complete Se precipitation was observed at all process conditions, while Te precipitation increased as the operating temperature was increased. Increasing the reagent quantity and temperature did, however, also result in increased copper and nickel co-precipitation.Regression models were used to perform numerical analyses to determine suitable operating conditions. Predictions with this numerical approach suggested that precipitation with 200% excess thio-urea at a temperature of 80 °C and a pressure of 7 bar would yield 98% Rh, 75% Ru, and 48% Ir precipitation with less than 5% Cu and Ni co-precipitation; these results could be experimentally validated.     

Selective leaching of penalty elements from copper concentrates: A review

Custom copper smelters impose substantial financial penalties for the presence of deleterious impurity elements in copper concentrates and can outright reject concentrates which contain impurity elements in concentrations that exceed specified values. Hence, there is strong motivation to remove penalised impurity elements from copper concentrates at the mine site before shipping to custom smelters. A number of leach systems have been developed for the selective extraction of penalty elements from copper concentrates, including: alkaline sulphide leaching (ASL); hypochlorite leaching; dilute sulphuric acid leaching with aluminium sulphate; and combined pressure oxidation (POX) leaching with copper precipitation leaching. This paper reviews these four systems with emphasis on the leaching behaviour of penalty elements. ASL has previously been employed in industry for the selective extraction of As and Sb from tetrahedrite-rich copper concentrates. Sodium sulphide solution leaches As, Sb, and Hg from a large range of minerals, however, does not leach arsenopyrite, a mineral which often contains a significant portion of the total As in copper concentrates. Hypochlorite leaching extracts As associated with enargite minerals. This leach system benefits from superior rates of As extraction when compared with ASL, and for this reason, has gained recent interest within the research community. Two major issues have been identified with hypochlorite leaching of copper concentrates. These are poor reagent selectivity towards As-bearing minerals and high levels of hypochlorite consumption. Unless these two issues are resolved it is unlikely that hypochlorite leaching will be employed in commercial processes. Dilute sulphuric acid leaching with aluminium sulphate is used to extract F associated with fluorite. This leach system also extracts F associated with apatite and chlorite. Laboratory-scale experiments and extensive operating experience have indicated that fluorite can be substantially leached from copper concentrates without addition of aluminium sulphate provided that the concentration of sulphuric acid in the leach solution is sufficiently high (at least 40 g L⁻¹). POX/copper precipitation leach systems have potential to extract a large number of penalty elements from copper sulphide concentrates while simultaneously upgrading the concentration of copper in the concentrate. Two patented POX/copper precipitation leach processes have been specifically developed for the deportment of penalty elements. These two processes are reviewed in detail.     

Electro-remediation of copper mine tailings. Comparing copper removal efficiencies for two tailings of different age

This work compares and evaluates the copper removal efficiency when applying electric fields to two mine tailings originating from the same mine but of different age. Eight experiments were carried out – four on tailings deposited more than 20 years ago (old tailings) and four on tailings deposited less than 2 years ago (new tailings). Parameters analyzed were the applied voltage drop, acid concentration during pretreatment, and the use of either passive or ion exchange membranes in the experimental setup.The comparison of the results confirms that there are differences in the electroremediation between the two tailings, even if the pH is similar and a mineralogical analysis showed similarities between the samples with respect to composition. It was found that an electroremediation is more favorable on the old tailings. The results showed that the best experimental conditions for both tailings is a pretreatment with H₂SO₄ 1 M followed by applying 40 V for 7 days, using ion exchange membranes. In this case 16.7% of copper was removed from the anode section for the old tailings, whereas only 11.2% was removed from the new tailings. The current efficiencies with respect to copper for the old and new tailings were 1.7% and 1.6%, respectively.     

Post-regrind selective depression of pyrite in pyritic copper–gold flotation using aeration and diethylenetriamine

The present study investigates the effect of aeration and diethylenetriamine (DETA) on the selective depression of pyrite in a porphyry copper–gold ore, after regrinding (at grind sizes, d₈₀ = 38 and 8 μm) with respect to Au recovery and grade using oxygen demand tests, flotation, QEMSCAN, X-ray spectroscopy (XPS) and EDTA extraction analysis. It was found that pyrite depression increases after aeration and with decreasing grind size. This was observed to be due to the markedly higher oxygen consumption rate of pyrite at the 8 μm (kla = 0.10 min⁻¹) than at the 38 μm grind size (kla = 0.02 min⁻¹). The addition of DETA improved pyrite depression (9% with aeration only versus 39% with aeration + DETA) at the 38 μm grind size. Gold and copper flotation recovery followed pyrite recovery for the two grind sizes using XD5002 in the presence of air and DETA.The surface analysis (XPS and EDTA extraction) revealed that the significant pyrite depression at the 8 μm grind size was due to increased amount of surface iron oxides, oxy-hydroxides (FeO/OH), sulphate species and increased liberation of mineral phases (QEMSCAN analysis), whilst the poorer pyrite depression at the 38 μm grind size was due to insufficient liberation of mineral phases and the persistence of activating Cu on the pyrite surface. The addition of DETA increased pyrite depression at the coarser grind size due to a significant reduction in Cu(I)S and increased Cu(II)O species, correlating with the flotation results of pyrite under this test condition. Two-stage copper and pyrite flotation, followed by Au cleaning after regrinding to 38 μm grind size, under high pH or aerated condition is proposed as the recommended route to optimise Au flotation.     

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.