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.     

Dissolution and passivation mechanisms of chalcopyrite during bioleaching: DFT calculation,XPS and electrochemistry analysis

In this work, density functional theory (DFT) calculation, X-ray photoelectron spectroscopy (XPS) and electrochemistry analysis were carried out to investigate the dissolution process and passivation mechanisms of chalcopyrite in the presence of sulfur and iron oxidizing microorganisms. Both DFT calculation and XPS analysis indicated that the formula of chalcopyrite should be Cu + Fe³ + (S²⁻)₂. Disulfide (S₂²⁻) and polysulfide (S_n²⁻) can be easily formed on the surface of chalcopyrite due to the surface reconstruction. The dissolution process of chalcopyrite in bioleaching was mainly dependent on redox potential. Chalcopyrite was predominantly directly oxidized to polysulfide when redox potential was lower than about 350 mV vs. Ag/AgCl and resulted in low dissolution rate. When redox potential was in the range of about 350–480 mV vs. Ag/AgCl, chalcopyrite was mainly transformed to intermediate species of Cu₂S rather than polysulfide, thus resulting in high dissolution rate. When redox potential was higher than about 480 mV vs. Ag/AgCl, chalcopyrite was principally directly oxidized to polysulfide which caused the passivation of chalcopyrite. Finally, a model of dissolution and passivation mechanisms of chalcopyrite in the presence of sulfur and iron oxidizing microorganisms was provided.     

Effects of quartz addition on chalcopyrite bioleaching in shaking flasks

This paper studies the effects of quartz on bioleaching of chalcopyrite by Acidithiobacillus ferrooxidans, LD-1 through shaking flask experiments. The results showed that quartz concentration can affect the copper extraction. After 32 days, copper extraction of the leaching system at 50 g L⁻¹ quartz concentration increased by about 20%, compared with that of the leaching system without quartz. XRD analysis showed that the amounts of jarosite on the chalcopyrite surface may reduce by the mechanical friction action between fine particles of quartz and chalcopyrite. The analysis of SEM indicated that the surfaces of chalcopyrite particles were eroded by different degrees and the degrees of change were the same as the effects of quartz concentration on copper extraction.     

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.     

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.     

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%.     

Possibilities for Co(III) dissolution from an oxidized ore through simultaneous bioleaching of pyrite

Solubilisation of Co(III) from a heterogenite met in copper cobaltiferous oxide ore has been realized through reductive leaching using ferrous iron generated via bio-oxidation of pyrite. Biotic and abiotic experiments at various pulp densities and redox potentials have been performed and results compared. Cobalt leaching at elevated redox potential is possible, offering cost reduction benefits due to reduced consumption of ferrous iron. At elevated potential of 625 mV, however, the initial rate of cobalt leaching has been found as 115 mg/(g ore)⋅(24 h), lower than the rate of 865 mg/(g ore)⋅(24 h) registered at 505 mV. Less stochiometric amount of ferrous iron was required when cobalt leaching was coupled to pyrite bioleaching, with 75% of cobalt recovered for 12 h at the optimally found conditions. It could be inferred that the Fe³⁺–Fe²⁺ cycle exists and is efficiently maintained through bacterial presence in the studied system.     

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.     

Bioleaching of chalcopyrite by Acidithiobacillus ferrooxidans

Acidithiobacillus ferrooxidans (A. ferrooxidans) was selected to experimentally study the effects of bacteria on the oxidation of chalcopyrite. The results indicated that A. ferrooxidans remarkably promoted the oxidation of chalcopyrite. The pH of the cell broth medium was observed to increase and then decrease during the bioleaching experiment. The number of suspended bacteria in the bio-oxidation process could be divided into three stages: the initial 4 days, in which the bacteria attached to the chalcopyrite surface and the number of suspended bacteria slightly decreased; day 5 to day 52, in which the suspended bacteria clearly increased with time and reached a maximum of 3.58 × 10⁷ cells/L on day 52; and day 53 to day 80, in which the number of suspended bacteria decreased. Other parameters such as redox potential (Eh) and iron ion concentrations increased with time. SEM micrographs showed that the cells were directly attached to the erosion pits on the smooth surfaces of the chalcopyrite. The erosion pits were similar to the bacteria in shape and size, and thus, the pits were likely products of dissolution by organic acids secreted by the attached cells. Compared to the unoxidized chalcopyrite, the elemental sulfur of the eroded chalcopyrite was clearly reduced, and the elemental oxygen was slightly increased. Moreover, a biofilm was present on the surfaces of the chalcopyrite particles. Therefore, the adherence of the cells to the mineral surfaces played a predominant role in altering the mineral appearance, which is important during the leaching of chalcopyrite.     

Cooperative bioleaching of chalcopyrite and silver-bearing tailing by mixed moderately thermophilic culture: An emphasis on the chalcopyrite dissolution with XPS and electrochemical analysis

In this work, the interactions between one sample of silver-bearing tailing (223 g/t silver) and chalcopyrite during bioleaching by mixed moderately thermophilic culture were investigated. Bioleaching results showed that copper can be almost totally extracted from chalcopyrite as the result of addition of the silver-bearing tailing, and silver (Ag) extraction can be significantly improved with the addition of chalcopyrite. Hence, cooperative bioleaching process of chalcopyrite and the chosen silver-bearing tailing was feasible. Ag mainly occurred as silver sulfate (Ag₂SO₄), and further work of enhancing the Ag extraction and its recovery is currently in progress. The catalytic effect of the silver-bearing tailing on chalcopyrite dissolution was investigated mainly with X-ray photoelectron spectroscopy (XPS) and electrochemical analysis. Results proved that the presence of the silver-bearing tailing enhanced the oxidation rate of chalcopyrite and also eliminated the passivation effect of polysulfide, thus resulting in an extremely high copper extraction.     

Mathematical modeling of thermophilic bioleaching of chalcopyrite

Previous studies have shown that the different preferences of thermophiles to oxidize S⁰ or Fe²⁺ is reflected by different [Fe³⁺]/[Fe²⁺] levels in solution. In those studies it was concluded that [Fe³⁺]/[Fe²⁺] governs the thermophilic bioleaching of chalcopyrite rather than temperature or pH. Therefore, the proposed model is mainly based on the finding that thermophilic bioleaching of chalcopyrite is governed by [Fe³⁺]/[Fe²⁺] that result from the activity of thermophiles. A direct interaction between chalcopyrite and thermophiles is neglected because it has been reported that this is not a general behavior for all thermophiles. The case of constant temperature, initial pH 1.5–2.5, and chalcopyrite concentrates is considered. The main assumption is that chalcopyrite can be anodically oxidized or cathodically reduced depending on [Fe³⁺]/[Fe²⁺] in solution. When chalcopyrite is oxidized at high [Fe³⁺]/[Fe²⁺] levels, Cu²⁺ is formed directly at low rates: CuFeS₂ + 4Fe³⁺ → Cu²⁺ + 5Fe²⁺ + S⁰. Whereas, when chalcopyrite is reduced at low [Fe³⁺]/[Fe²⁺] levels, an intermediate (Cu₂S) is formed at higher rates: CuFeS₂ + Fe²⁺ + Cu²⁺ + 2H⁺ → Cu₂S + 2Fe³⁺ + H₂S. Because the oxidation of Cu₂S is relatively fast: Cu₂S + 4Fe³⁺ → 2Cu²⁺ + S⁰ + 4Fe²⁺, its accumulation is assumed to be negligible. To take into account the possibility of chalcopyrite being oxidized or reduced depending on [Fe³⁺]/[Fe²⁺] in solution, the principle of mixed potentials is used. The model is validated by comparing the calculated and measured values of copper extraction, total iron in solution, and pH.     

Synchrotron X-ray photoelectron spectroscopic study of the chalcopyrite leached by moderate thermophiles and mesophiles

SXPS (Synchrotron X-ray Photoelectron Spectroscopy) and NEXAFS (Near Edge X-ray Absorption Fine Structures) have been applied to study the surface chemical species of chalcopyrite leached by a moderate thermophilic consortia, Leptospirillum ferrooxidans and a mesophilic mixed culture of L. ferrooxidans, Acidithiobacillus ferrooxidans and Acidithiobacillus thiooxidans. A sulfur-rich layer dominated by S_n²⁻ developed with time, which was found to control the rate of bioleaching. Fe L_2,3-edge NEXAFS and Fe 2p spectra indicate the formation of jarosite during bioleaching. Thermophiles significantly enhanced the leaching efficiency, in which 1.34 g/L copper was dissolved in 25 days, while less than 0.3 g/L copper was released in 30 °C bioleaching. This was mostly caused by the increased abiotic reaction rate. The solution copper concentration in presence of L. ferrooxidans was higher than that with mesophilic mixed culture, which suggests the synergistic effect of mixed microorganisms did not play a comparably important role as temperature under the conditions used in this study. Explicit evidence of elemental sulfur was only found in samples leached by L. ferrooxidans by Raman spectroscopy. However, the formation of elemental sulfur does not significantly hinder the leach rate.     

Why Zijinshan copper bioheapleaching plant works efficiently at low microbial activity – Study on leaching kinetics of copper sulfides and its implications

Zijinshan commercial bioheapleaching–SX–EW plant has been in operation since December 2005. 13.9 million tonnes of low grade copper sulfide ore (Cu 0.38%) has been leached efficiently with an annual recovery of 80% in this distinct leaching system with high iron concentration, high temperature, high acidity and low microbial activity. This paper is to study the key factors which influence the heap operation in Zijinshan. Dissolution kinetics of diginite and covellite was carried out in acid ferric sulfate solution under controlled redox potential. Moreover, column bioleaching tests were conducted in the laboratory with raffinate from the plant. Both results showed that the main factors accelerating dissolution rate of Cu_1.8S/CuS were mainly temperature and ferric concentration, while redox potential only had a slight positive effect. The present work provides an insight into mineral kinetics to understand this unique bioheapleaching operation and gives further information for optimization of the existing practices.     

Surface characterisation,collector adsorption and flotation response of enargite in a redox potential controlled environment

We previously investigated oxidation of the surface of natural enargite (Cu₃AsS₄) under potentiostatic control and the formation of oxidation species at the mineral surface at selected applied potentials in the oxidative range. Here we further extended the research by incorporating flotation collectors into the system. Electrochemical techniques, X-ray photoelectron spectroscopy (XPS) and microflotation in a redox potential controlled environment were applied to examine surface properties, collector adsorption and flotation response of enargite in pH 10 solutions of sodium ethyl xanthate (SEX) and sodium dialkyl dithiophosphinate (3418A). The spectral details of XPS analysis of electrochemically treated enargite surfaces show significant adsorption of SEX and 3418A collectors onto enargite at an applied voltage of +516 mV, but no adsorption of both collectors at −400 mV. The results of XPS analysis agree with the floatability of enargite determined by microflotation, showing that the flotation recovery was highest at high oxidative potential (+516 mV), then decreased at low oxidative potential (+100 mV) and was very poor at −400 mV. These results confirm that enargite floatability can be efficiently controlled electrochemically.     

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.     

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.     

Bioleaching of six nickel sulphide ores with differing mineralogies in stirred-tank reactors at 30 °C

A bioleaching study was conducted with six nickel sulphide ores from different geographical locations across Canada. Mineralogical and chemical examination revealed considerable variability between the samples, particularly in the silicate phases. The ores contain 0.3–1% nickel, primarily in pentlandite and secondarily in pyrrhotite. Copper is present primarily in chalcopyrite, and cobalt in pentlandite. The ores were subjected to the same crushing and grinding procedure, and bioleached under the same conditions for 3 weeks with a mixed culture of iron- and sulphur-oxidizing bacteria. Stirred-tank experiments with finely ground ore (−147 μm) at 30 °C were conducted to assess the effect of pH (2–5) and the impact of the bacteria. Nickel extraction from pentlandite and pyrrhotite during bioleaching at pH 2 and 3 was generally good (49–86% after 3 weeks), and cobalt extraction tracked nickel extraction over most conditions. All six ores showed a similar response to a change in pH; an increase in pH from 2 to 3 resulted in approximately the same nickel and cobalt extraction (within statistical error), and a statistically significant reduction in sulphuric acid consumption, dissolved iron, and magnesium extraction.     

The influence of pyrite content on the flotation of chalcopyrite/pyrite mixtures

In the flotation of copper ores, several processing plants report that copper recovery is affected by the proportion and reactivity of pyrite in the ore, with the effect becoming more intense when the feed particles are finer as a result of regrinding. In this work, a mixed model mineral system consisting of chalcopyrite (CuFeS₂) and pyrite (FeS₂) with varying pyrite content (20–80 wt.%) was used to investigate the effect of pyrite on the pulp chemistry and chalcopyrite recovery. Flotation tests showed that chalcopyrite flotation rate, recovery and grade, as well as the pulp oxidation potential, decreased with increasing pyrite content whilst pyrite recovery increased. Surface analysis (XPS, ToF-SIMS and EDTA) indicated that copper activation of pyrite increased with increasing pyrite content, facilitating pyrite recovery. The decrease in chalcopyrite recovery can be attributed to increased surface oxidation.     

Passivation of chalcopyrite during its chemical leaching with ferric ion at 68 °C

The passivation of chalcopyrite in the presence of ferric sulphate solutions was investigated at 68 °C. The effect of different variables (pulp density, pH and the presence of oxygen) on both the copper dissolution rate and the formation of solid compounds was studied. The leaching tests were carried out in stirred flasks at 180 rpm with 100 mL of a Fe³⁺/Fe²⁺ sulphate solution, varying the pulp density between 0.1 and 5% and the pH between 0.5 and 2.0 and both in aerobic and anaerobic conditions. Ferric ion seems to be responsible for the oxidation of chalcopyrite but also promotes its passivation. Ferrous ion plays a key role in the process by controlling the nucleation and precipitation of jarosites, which finally cause the passivation of chalcopyrite.     

Selective flotation of cassiterite with benzohydroxamic acid

The flotation of cassiterite mineral from gangue with a collector benzohydroxamic acid (BHA), and the interactions between the BHA and cassiterite have been investigated. It is shown through microflotation that the BHA is able to flot cassiterite very well, calcite quite limitedly, and quartz not at all, so the selective separation of cassiterite–quartz mixture was readily achieved; while for the efficient separation of cassiterite–calcite mixture containing 48.94% SnO₂, sodium hexametaphosphate (SHMP) was needed as a depressant for the gangue, and under the condition of the BHA 100 mg L⁻¹, SHMP 3.5 mg L⁻¹, a cassiterite concentrate with the grade of 85.50% SnO₂ was obtained with the recovery of SnO₂ 95.5%. Batch flotation further demonstrated that for an industrial tin slime, which contained 0.42% Sn, 13.65% SiO₂, 24.14% CaO, 16.60% MgO, 4.50% Al₂O₃ and 6.58% Fe, the tin recovery of 84.5% after one separation was reached with the concentrate grade of 1.84% Sn under the condition of the BHA 178 mg L⁻¹, SHMP 27 mg L⁻¹. In terms of zeta potential and infrared spectra studies the main interactions between the collector BHA and the mineral cassiterite in a flotation system are chemisorption with the formation of Sn–BHA compounds rather than electrostatic attractions between them.