Biohydrometallurgy in Turkish gold mining: First shake flask and bioreactor studies

The first laboratory tests on biooxidation and cyanidation of gold ores in Turkey were carried out using samples of the Copler Gold Mine. Over a 3 year R&D test period, mixed bacterial/archaeal cultures improved biooxidation of the Copler ore. The highest sulphide oxidation of 87.35% over 432 h was achieved in shake flasks in the presence of the mixed culture (MODM: Sulfolobus acidophilus and Sulfolobus thermosulfidooxidans). Bioreactor tests resulted in greater dissolution rates for iron and arsenic than did shake-flask tests, which led to a greater extent of sulphide oxidation within a shorter period of time. The maximum sulphide oxidation in the bioreactor tests was 97.79% after 240 h when the EXTM (Acidianus brierleyi and Sulfolobus metallicus) mixed culture was used. After the biooxidation experiments with solids contents of 10% and 20% (w/v), the gold recovery from the oxidised ore was lower than that achieved in the presence of 5% solids (w/v) because the extent of sulphide oxidation was reduced as the pulp density increased. A strong correlation between the sulphide oxidation and gold recovery was also established. The highest gold recovery of 94.48% was achieved during cyanidation from the biooxidised ore produced from the experiment conducted using the EXTM mixed culture.     

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.     

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.     

Alkaline leaching of zinc from electric arc furnace steel dust

Electric arc furnaces (EAF) generate about 10–20 kg of dust per metric ton of steel, which constitute a hazardous waste, known as EAF dust. This dust contains a remarkable amount of non-ferrous metals, which include zinc, cadmium, lead, chromium and nickel that could be recovered, reducing the environmental impact of the leachable toxic metals, and generating revenue. In this paper, different alkaline leaching techniques were tested in order to dissolve the zinc present in an EAF dust: (i) conventional agitation leaching; (ii) pressure leaching; (iii) conventional leaching following a microwave pretreatment and (iv) leaching with agitation provided by an ultra-sonic probe. Temperature and sodium hydroxide concentration were the variables tested. The highest zinc recovery from the EAF dust, containing about 12% of zinc, was about 74%. This was achieved after 4 h of leaching in a temperature of 90 °C and with a sodium hydroxide concentration of 6 M of the leaching agent.     

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).     

Density functional theory study of α-Bromolauric acid adsorption on the α-quartz (1 0 1) surface

Adsorption mechanism of collector α-Bromolauric acid (CH₃(CH₂)₉CHBrCOOH, α-BLA) on α-quartz (1 0 1) surface has been investigated by first-principles calculations based on density functional theory (DFT). The interaction energies of H₂O molecule, calcium ions (Ca²⁺), hydroxyl ions (OH⁻), calcium hydroxyl ions (Ca(OH)⁺), and α-BLA⁻ ions with α-quartz (1 0 1) surface were in the order of Ca(OH)⁺ < Ca²⁺ < OH⁻ < H₂O < α-BLA⁻. The results revealed that the collector α-BLA cannot adsorb on α-quartz (1 0 1) surface due to the hindrance of hydration shell of quartz surface, while Ca(OH)⁺ could repulse the hydration shell and consequently adsorb on quartz surface, which further leads to the adsorption of the collector α-BLA⁻ anions on Ca(OH)⁺-activated quartz surface. Mulliken populations analysis of the external oxygen atom (O2) of quartz surface, calcium atom (Ca) of Ca(OH)⁺, and oxygen atom (O1) of collector α-BLA⁻ (–OH group) shows that the electron transfer between the Ca–O1 and Ca–O2 atoms. The overlap area of electron density between Ca–O1 and Ca–O2 atoms indicates strong interactions among the three atoms of Ca, O1, and O2, suggesting that Ca(OH)⁺ ions act as a bridge between the α-quartz (1 0 1) surface and the α-BLA collector.     

Bioleaching and chemical leaching as an integrated process in the zinc industry

This work sought to integrate bioleaching and chemical leaching as a cost-effective process to treat zinc sulphides. The continuous bioleaching of a sphalerite concentrate, assaying 51.4% Zn, 1.9% Pb, 31.8% S and 9.0% Fe with mesophile iron and sulphur-oxidizing bacteria followed by chemical leaching of the bioleaching residue were assessed. In the bioleaching step, the first reactor was used to produce Fe(III) concentrations as high as 20 g/L. This solution was fed to the subsequent bioleaching reactors to oxidize sphalerite. It was possible to achieve 30% zinc extraction for 70 h residence time. In chemical leaching experiments, carried out with the residue of the bioleaching step, the effects Fe_total and acidity on zinc extraction were studied. It was noticed that Fe(III) concentrations over 12 g/L did not affect zinc recoveries. Furthermore, the higher the acidity, the larger the zinc recovery, for experiments carried out up to 181 g/L sulphuric acid. The results have demonstrated that it is possible to devise a new process capable of achieving 96% zinc extraction, similarly to the conventional roasting–leaching–electrolysis process.     

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.     

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 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⁺).     

Cyanide oxidation by ozone in a steady-state flow bubble column

The treatment of the effluent produced in the cyanidation of gold and silver ores is one of the main problems that the precious metals industry currently faces. These effluents contain variable amounts of free and complex cyanide (stable and unstable). Different methods for the cyanide elimination have been used; nevertheless, most of the times the consumption of reagents increases operating costs to prohibiting levels. As well, the process may give rise to by-products that are also toxic.Ozone oxidation is a promising alternative for cyanide control that offers several advantages. In this work, the oxidation of cyanide in synthetic alkaline solutions by ozone was studied in a countercurrent bubble column. Dissolved ozone, cyanide and cyanate profiles were obtained along the column, demonstrating that the cyanide/ozone reaction may take place all along the reactor. This is found to occur when the mole of O₃/mole of CN rate (specific ozone dose) is higher than one.The oxidation efficiency depends mainly on the specific ozone dose (moles of ozone fed per mole of cyanide), leading to 90% efficiency when the specific ozone dose is 1.2 mol O₃/mol CN or higher. The ozone consumption obtained in these tests (steady-state flow pilot-scale column) was similar to that obtained at the laboratory scale in a semi-batch mechanically stirred reactor (1.2–1.4 mol O₃/mol CN), suggesting that the size of the reactor has no effect on ozone consumption.     

Uranium recovery from strong acidic solutions by solvent extraction with Cyanex 923 and a modifier

Uranium stripping with strong acid solution is always highly desired due to its simple operation and less pollution. However, intensive acid neutralisation for uranium precipitation in the subsequent step limited its application. A new solvent extraction process has been developed to transfer uranium from strong to weak sulphuric acid solutions suitable for uranium precipitation without intensive neutralisation. An organic system consisting of 10% Cyanex 923 and 10% isodecanol as the modifier in ShellSol D70 was optimised for the process. It was found that uranium was extracted efficiently from 4 to 6 M H₂SO₄ solutions with the organic system, and it could be efficiently stripped with 0.2–0.5 M H₂SO₄ solutions. Both extraction and stripping kinetics of uranium were very fast, reaching the equilibrium within 0.5 min. Temperature between 30 and 60 °C has slight effect on uranium extraction and stripping. Four theoretical stages could effectively extract more than 98% uranium from a solution containing 17.5 g/L U and 6.0 M H₂SO₄ at an A/O ratio of 1:1.5, and it could generate a loaded organic solution containing about 12 g/L U. More than 99% U could be stripped from the loaded organic solution containing 14.6 g/L U with 0.5 M H₂SO₄ using five stages at an A/O ratio of 1:3. As a result, the loaded strip liquor containing more than 40 g/L U would be obtained which is suitable for uranium recovery by precipitation using hydrogen peroxide. A conceptual process has been proposed for uranium transfer from strong to weak sulphuric acid solutions for its recovery.     

Statistical analysis of bioleaching copper,cobalt and nickel from polymetalic concentrate originating from Kamoya deposit in the Democratic Republic of Congo

The effects of five parameters, temperature, pH, leaching duration, stirring speed and pulp density on the bioleaching of copper, cobalt and nickel from a polymetallic flotation concentrate were investigated. The leaching was carried out according to the L₂₅ (5⁵) orthogonal design. The optimal values of the parameters were determined using a Taguchi method through signal-to-noise analysis. ANOVA was applied to verify the individual contribution of each parameter and their degree of significance. It was found out that pulp density was the most influential factor on the bioleaching yield of the three metals altogether, followed by pH and temperature. For the copper bioleach, the following optimal parameters were determined: temperature – 37.5 °C, pH 1.6, leaching duration – 20 days, stirring speed – 350 rpm and pulp density – 7.5%. Verification experiments conducted according to these optimal parameters brought copper yield of 72.6%. For the cobalt bioleach, SEM observations of pure carrolite indicated a progressive bacterial colonization of mineral surface with time.     

Cooperative effect of chalcopyrite and bornite interactions during bioleaching by mixed moderately thermophilic culture

The cooperative interactions between chalcopyrite and bornite during bioleaching by mixed moderately thermophilic culture were investigated mainly by bioleaching experiments and electrochemical experiments. Bioleaching results showed that a cooperative effect existed between chalcopyrite and bornite. When the mass ratio of chalcopyrite to bornite was 3:1, an extremely high copper extraction of more than 88% was achieved after bioleaching for 27 days. One of the major reasons for the cooperative effect was that a certain redox potential range (370–450 mV vs. Ag/AgCl) could be maintained for a long period of time during bioleaching due to the mixture of chalcopyrite and bornite. Electrochemical measurements revealed that chalcopyrite was much easier to be reduced than oxidized, while bornite was prone to be directly oxidized. Hence, galvanic effect between chalcopyrite and bornite enhanced the reduction of chalcopyrite to secondary copper-iron species and promoted the oxidative dissolution of bornite. Therefore, redox potential controlling and galvanic effect both contributed to the cooperative bioleaching of chalcopyrite and bornite.     

Engineered strains enhance gold biorecovery from electronic scrap

Gold bioleaching from electronic scrap materials (ESM) was examined using the bacterium Chromobacterium violaceum which produces and detoxifies cyanide, one of the few lixiviants capable of leaching gold. Gold recovery by the wild-type C. violaceum and two genetically engineered strains (pBAD and pTAC) with an additional cyanide-producing operon were investigated and compared. The ESM was pretreated to remove metals competing for metal cyanide complexation with gold. The effect of pulp density on leaching performance by the various strains was also investigated. The pBAD strain produced the highest cyanide concentration, and achieved the highest gold recovery of 30% at 0.5% w/v pulp density, compared to 11% recovery by the wild-type bacteria. Our results demonstrated the application of lixiviant metabolic engineering in the construction of enhanced bioleaching microbes for the bioleaching of precious metals from electronic waste.     

Degradation of thiocyanate in aqueous solution by persulfate activated ferric ion

Thiocyanate formation from cyanidation of gold bearing ores is becoming a more common problem during gold processing. In this work, the application of an advanced oxidation process based on the use of persulfate (S₂O₈²⁻) as an environmentally friendly oxidant in the presence of ferric ion for destruction of a persistent and non-volatile inorganic contaminant, such as thiocyanate, in aqueous solutions is reported for the first time. The influence of various reaction parameters like ferric ion and persulfate dosage, initial thiocyanate concentration and the influence of radical scavenger are examined. An accelerated reaction using S₂O₈²⁻ to destroy thiocyanate can be achieved via chemical activation with Fe³⁺ to generate highly reactive sulfate anion-radicals (SO₄⁻). The results showed that degradation efficiency was negligible when persulfate was used alone, ferric ions significantly improved the degradation efficiency of thiocyanate at ambient temperature. Under the optimum molar ratios ([S₂O₈²⁻]:[SCN⁻] = 5:1 and [S₂O₈²⁻]:[Fe³⁺] = 1:0.2), 99% of thiocyanate present in aqueous solution at the initial concentration range of 1.72–17.2 mM was degraded within 60 min of reaction time. To evaluate the contribution of reactive free radicals generated through Fe(III)-mediated activation of persulfate to thiocyanate degradation, quenching experiments using methanol as the radical quenching agent were carried out. The obvious decrease in thiocyanate oxidation efficiency in the presence of methanol confirmed that the radical-based pathway was the dominant mechanism in Fe³⁺/S₂O₈²⁻ system. The degradation of thiocyanate was accompanied by the formation of cyanide as the main final product of the reaction. Thus the catalytic oxidation of thiocyanate makes it possible to return NaCN into the production process for leaching of precious metals. The work presents an efficient and environmentally acceptable wastewater treatment process applicable in mining facilities utilizing cyanidation of sulfide ores and/or concentrates.     

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.     

Bioleaching of djurleite using Acidithiobacillus ferrooxidans

The bioleaching of djurleite using Acidithiobacillus ferrooxidans (LD-1) was investigated in this paper. Experiments were carried out in shake flasks at pH 2.0, 160 r/min and 30 °C. The leaching residues were analyzed using X-ray diffraction (XRD), Scanning electron microscopy (SEM) and X-ray photoelectron spectroscopy (XPS). The total copper extraction of djurleite under optimal condition reached 95.12%. The XRD analysis indicated the residues mainly consisted of ammoniojarosites and S₈. It was observed by the SEM image that the djurleite was heavily etched. The XPS results confirmed the intermediate product formed during djurleite leaching was CuS. The result indicates the reaction pathway is: Cu₃₁S₁₆ → CuS → tCu²⁺ and S⁰.     

Kinetics of high-sulphur and high-arsenic refractory gold concentrate oxidation by dilute nitric acid under mild conditions

On one hand, high-sulphur and high-arsenic refractory gold concentrate (HGC) leads to regional ecological damage. On the other hand, it contains a lot of valuable elements. So utilization of it can bring social and environmental benefits. In this paper, the kinetics of HGC oxidation by dilute nitric acid under mild conditions was investigated. The effects of particle size (50–335 μm), reaction temperature (25–85 °C), initial acid concentration (10–30 wt.%) and stirring speed (400–800 rpm) on the iron extraction rate (C_r) were determined. It is obvious that C_r increases with the rise of initial nitric acid concentration, reaction time and stirring speed, but decreases with the increase of particle size. Oxidation kinetics indicates that the rate of reaction is diffusion controlled. The activation energies were determined to be 10.70 kJ/mol in the 10% HNO₃ and 12.25 kJ/mol in the 25% HNO₃.     

The effect of flotation and parameters for bioleaching of printed circuit boards

Waste electrical and electronic equipment (WEEE) is currently one of the fastest growing waste streams in the world. Typical for WEEE is the high content of valuable and precious metals, as well as harmful contaminants like halogens, flame retardant chemicals and plastics. Currently, WEEE treatment and metal recovery methods are imperfect, polluting and energy intensive. In this paper, novel treatment possibilities are outlined for printed circuit boards (PCB) utilizing both the flotation separation technique and acid bioleaching. Flotation, conducted after crushing and sieving of PCB, produced two fractions: metal-rich concentrate, which is more suitable for pyrometallurgical treatment than untreated PCB, and metal-poor froth suitable for acid bioleaching. It was seen that especially low pH (1.6), high initial Fe²⁺ concentration (7.8 g/l) and low PCB froth concentration in the bioleaching solution (50 g/l) were beneficial for the rapid and selective dissolution of copper. With these parameters, 99% of copper was solubilized from PCB froth in bioreactor treatment, with Cu (6.8 g/l) and Fe (7.0 g/l) being the only major metallic elements in bioleaching solution.