Single and binary component sorption of the fission products Sr2+, Cs+ and Co2+ from aqueous solutions onto sulphate reducing bacteria

This study investigates the removal of the fission products Sr²⁺, Cs⁺ and Co²⁺ in single and binary metal solutions by a sulphate reducing bacteria (SRB) biomass. The effect of initial concentration and pH on the sorption kinetics of each metal was evaluated in single metal solutions. Binary component equilibrium sorption studies were performed to investigate the competitive binding behaviour of each metal in the presence of a secondary metal ion. Results obtained from single metal equilibrium sorption studies indicated that SRB have a higher binding capacity for Sr²⁺ (q_max = 416.7 mg g^?1), followed by Cs⁺ (q_max = 238.1 mg g^?1), and lastly Co²⁺ (q_max = 204.1 mg g^?1). Among the binary systems investigated, Co²⁺ uptake was the most sensitive, resulting in a 76% reduction of the sorption capacity (q_max) in the presence of Cs⁺. These findings are significant for future development of effective biological processes for radioactive waste management under realistic conditions.     

Leaching of chalcopyrite (CuFeS2) with an imidazolium-based ionic liquid in the presence of chloride

The effects of independent variables such as, temperature, concentration of ionic liquid (1-butyl-3-methyl-imidazolium hydrogen sulphate, [bmim][HSO₄]), chloride and sulphuric acid on copper extraction from chalcopyrite (CuFeS₂) ore were studied by surface optimization methodology. The Central Composite Face approach and a quadratic model were applied to the experimental design. The optimal copper extraction conditions given by the above methodology were 20% (v/v) of [bmim][HSO₄] in water, 100 g L⁻¹ chloride, and 90 °C. The concentration of chloride and the temperature together exert a synergistic effect in enhancing chalcopyrite dissolution. Experimental data were fitted by multiple regression analysis to a quadratic equation and analyzed statistically. A model was developed for predicting copper extraction from CuFeS₂ ore with variables such as Cl⁻, [bmim][HSO₄], H₂SO₄ concentrations and temperature in the range studied. The activation energy was calculated to be 60.4 kJ/mol (temperature range 30–90 °C), indicative of chemical control of the reaction and [bmim][HSO₄] acts as an acid in the reaction.     

Thorium and uranium extraction from rare earth elements in monazite sulfuric acid liquor through solvent extraction

This paper describes the process of extraction of thorium and uranium from the sulfuric liquor generated in the chemical monazite treatment through a solvent extraction technique. The influence of the extractant type and concentration, contact time between phases, type and concentration of the stripping solution and aqueous/organic volumetric ratio were investigated. The results indicated the possibility of extracting, simultaneously, thorium and uranium from a solvent containing a mixture of Primene JM-T and Alamine 336. The stripping was carried out with a hydrochloric acid solution. After selecting the best conditions for the process, a continuous experiment was carried out in a mixer-settler circuit using four stages in the extraction step, five stages of stripping and one stage of the solvent regeneration. A loaded stripping solution containing 29.3 g/L of ThO₂ and 1.27 g/L of U₃O₈ was obtained. The metals content in the raffinate was below 0.001 g/L, indicating a thorium extraction of over 99.9% and a uranium extraction of 99.4%. The rare earths content in the raffinate was 38 g/L of RE₂O₃.     

Oxidation of cyanide in effluents by Caro’s Acid

Caro’s Acid (peroxymonosulphuric acid: H₂SO₅) is a powerful liquid oxidant made from hydrogen peroxide that has been adopted for the detoxification of effluents containing cyanides in gold extraction plants in recent years.The present work reports the findings of a study on the kinetics of aqueous cyanide oxidation with Caro’s Acid. Experiments were conducted in batch mode using synthetic solutions of free cyanide. The experimental methodology employed involved a sequence of two 2³ factorial designs using three factors: initial [CN⁻]: 100–400 mg/L; H₂SO₅:CN⁻ molar ratio: 1–1.5–3–4.5; pH: 9–11; each one conducted at one level of Caro’s Acid strength which is obtained with the H₂SO₄:H₂O₂ molar ratio used in Caro’s Acid preparation of 3:1 and 1:1. The objective was the evaluation of the effect of those factors on the reaction kinetics at room temperature. Statistical analysis showed that the three investigated variables were found to be significant, with the variables which affected the most being the initial [CN⁻] and the H₂SO₅:CN⁻ molar ratio. The highest reaction rates were obtained for the following conditions: H₂SO₅:CN⁻ molar ratio = 4.5:1; pH = 9; and Caro’s Acid strength produced from the mixture of 3 mol of H₂SO₄ with 1 mol of H₂O₂. These conditions led to a reduction of [CN⁻] from an initial value of 400 mg/L to [CN⁻] = 1.0 mg/L after 10 min of batch reaction time at room temperature. An empirical kinetic model incorporating the weight of the contributions and the interrelation of the relevant process variables has been derived as: −d[CN⁻]/dt = k [CN⁻]^1.8 [H₂SO₅]^1.1 [H⁺]^0.06, with k = 3.8 (±2.7) × 10⁻⁶ L/mg min, at 25 °C.     

Ionic liquid-based modified cold-induced aggregation microextraction (M-CIAME) as a novel solvent extraction method for determination of gold in saline solutions

Modified-cold-induced aggregation microextraction (M-CIAME) was used for determination of gold in saline solutions. It is robust against the much higher concentration of salt (up to 40%). In this method sodium hexafluorophosphate (NaPF₆) was added to the sample solution containing Au-TMK complex and a very small amount of 1-hexyl-3-methylimidazolium tetrafluoroborate [Hmim][BF₄]. Afterward the solution was cooled in an ice bath and a cloudy solution was formed. After centrifuging, the extractant phase was analyzed using a spectrophotometric detection method. Under the optimum conditions, the limit of detection (LOD) was 0.7 ng mL^?1 and the relative standard deviation (RSD) was 1.65% for 50 ng mL^?1 gold. The method was applied for the determination of trace amount of Au in mineral and seawater with satisfactory results.     

Solvent extraction and separation of palladium(II) and platinum(IV) from hydrochloric acid medium with dibutyl sulfoxide

The solvent extraction and separation performances of Pd(II) and Pt(IV) from hydrochloric acid solutions were investigated using dibutyl sulfoxide (DBSO) diluted in kerosene. Pd(II) was strongly extracted by a lower concentration DBSO in a lower concentration hydrochloric acid solution while the reverse was obtained for Pt(IV) extraction. Based on independent extraction and separation experiments of Pd(II) and Pt(IV), the separation parameters of Pd(II) and Pt(IV), including dibutyl sulfoxide concentration, contact time of aqueous and organic phases, organic/aqueous (O/A) phase ratio and H⁺ concentration of aqueous phase, were studied in detail, and the optimal separation parameters were obtained and summarized as the following: dibutyl sulfoxide concentration 0.6–1.2 mol dm^?3, organic/aqueous (O/A) phase ratio 0.6–1.0, H⁺ concentration of aqueous phase 1.0–1.5 mol dm^?3 and contact time of two phases 5 min. The as-prepared separation parameters were corroborated by the extraction and separation from a synthetic stock solution containing Pd(II), Pt(IV) as well as several common impurities like Fe(II), Cu(II) and Ni(II). The results revealed that Pd(II) could be separated efficiently from Pt(IV) with a high separation coefficient of Pd(II) an Pt(IV) (2.7 × 10⁴) by predominantly controlling dibutyl sulfoxide and hydrochloric acid concentrations. The extraction saturation capacity of Pd(II) was determined from 1.0 mol dm^?3 HCl solution with 3 mol dm^?3 dibutyl sulfoxide and its experimental value exceeded 14 g dm^?3 under the experimental conditions.Stripping of Pd(II) from loaded organic phase was performed using a mixed aqueous solution containing NH₄Cl and ammonia solutes. Pd(II) (99.2%) was stripped using the stripping solution containing 3% (m/v) NH₄Cl and 5 mol dm^?3 ammonia, respectively.     

Extraction of titanium (IV) from acidic media by tri-n-butyl phosphate in kerosene

The extraction of titanium (IV) from sulfate, and nitrate solutions has been studied using tri-n-butyl phosphate (TBP) in kerosene. Extraction of titanium was affected by acid concentration over the range of 0.5–4 mol L^?1. The titanium distribution coefficient reached a minimum between 1 and 2 mol L^?1 acid for both sulfate and nitrate solutions. Third phase formation was observed in the extraction of titanium from acidic media at all condition tested. At the next stage, the stripping of titanium was studied using H₂SO₄, H₂SO₄ + H₂O₂ and Na₂CO₃. The kinetics of the stripping were very slow for H₂SO₄. The use of complex forming stripping agents (H₂SO₄ + H₂O₂) and Na₂CO₃ significantly improved the kinetics of stripping. About 98% recovery was achieved by extracting titanium from an aqueous nitrate solution using TBP and stripping with sodium carbonate.     

Simultaneous oxidation and immobilization of arsenite from refinery waste water by thermoacidophilic iron-oxidizing archaeon,Acidianus brierleyi

The use of a thermophilic acidophilic iron-oxidizing archaeon, Acidianus brierleyi, was investigated for oxidation and immobilization of As(III) from acidic refinery waste water. Some As(III) oxidation was measured in all Ac. brierleyi cultures independently of the presence or concentration of Fe(II) in bulk solution; the exception was at initial Fe(II) concentration ([Fe(II)]_ini) of 1000 mg l⁻¹ where As(III) oxidation became markedly facilitated and consequently approximately 70% of As was immobilized as amorphous ferric arsenate. Providing 1000 mg l⁻¹ Fe(III) instead of Fe(II) did not show the same effect, implying the importance of Fe(III) be microbially-produced and complexed in the archaeal EPS (extracellular polymeric substances) region for effective As(III) oxidation. The reaction towards secondary mineral formation shifted from ferric arsenate to jarosite at [Fe(II)]_ini of >1000 mg l⁻¹. Furthermore addition of jarosite seed crystals retarded the As(III) oxidation rate at [Fe(II)]_ini of 1000 mg l⁻¹. The observations indicate that by setting the appropriate bulk Fe(II)/As(III) ratio in Ac. brierleyi culture to achieve a certain concentration of Fe(III) within the EPS region, but at the same time to avoid jarosite formation, it is possible to maximize the As(III) oxidation rate and thus As immobilization efficiency. This study describes for the first time microbially-mediated simultaneous oxidation and immobilization of As(III) as ferric arsenate, using a thermoacidophilic iron-oxidizing archaeon, Ac. brierleyi.     

The effect of [Fe]TOT on the dissolution of synthetic Pb-doped UO2 and Th-doped UO2

The dissolution of synthetic Pb-doped UO₂ and Th-doped UO₂ was systematically studied to determine the influence of leach parameters [Fe]_TOT and ORP under standard leach conditions of: T = 50 °C, [H₂SO₄] = 15 g/L (0.15 M), and UO₂ = 100 mg/L. Results demonstrated reduced uranium dissolution in both systems compared to pure UO₂. This effect was greatest for Th-doped UO₂. The decrease in uranium dissolution between the doped systems and pure UO₂ was attributed to the formation of precipitate layers at the surface of the solid, slowing down or blocking uranium release. In the case of Pb-doped UO₂, the formation of a Pb sulphate phase was directly detected but in the case of Th-doped UO₂, no layer was found. For the latter system it was postulated that passivation of the Th-doped UO₂ surface occurs due to the formation of oxidised Th-rich phases Th(OH)₄, ThO₂ and ThO₂·nH₂O at the surface of grains preventing uranium release. In tests varying the ORP, there was an approximately linear dependence of the dissolution rate on [Fe]_TOT for both systems however the rate orders indicated a step change between an ORP of 420 and 460 mV. The specific influence of Fe^II showed that both Pb–UO₂ and Th–UO₂ exhibited two distinct regions of dissolution rate dependency similar to that previously noted for pure UO₂.     

Heavy metals removal from solution by palygorskite clay

The possible use of palygorskite clay, mined in the Dwaalboom area of the Northern Province of South Africa, as an adsorbent for the removal of metal ions such as lead, nickel, chromium and copper from aqueous solution, was investigated. In this work, adsorption of these metals onto palygorskite has been studied by using a batch method at room temperature. The results of adsorption were fitted to both the Langmuir and Freundlich models. Satisfactory agreement between experimental data and the model-predicted values was expressed by the correlation coefficient (R²). The Langmuir model represented the sorption process better than the Freundlich one, with correlation coefficient (R²) values ranging from 0.953 to 0.994. The adsorption capacity (Q₀) calculated from the Langmuir isotherm was 62.1 mg Pb(II) g⁻¹, 33.4 mg Ni(II) g⁻¹, 58.5 mg Cr(VI) g⁻¹ and 30.7 mg Cu(II) g⁻¹ at a pH of 7.0 at 25 ± 1 °C for a clay particle size of 125 μm. Kinetic investigations were performed to investigate the rate of adsorption of metal ions. The Lagergren’s first-order rate constants were calculated for different initial concentrations of metal ions. In batch mode adsorption studies, removal increased with an increase of contact time, adsorbent amount and solution pH. Adsorption of metals from the single-metal solutions was in the order: Pb > Cr > Ni > Cu. Data from this study proved that metal cations from aqueous solution can be adsorbed successfully in significant amounts by palygorskite. This opens up new possibilities and potential commercial uses in the palygorskite market.     

Biosorptive removal of Cd and Zn from liquid streams with a Rhodococcus opacus strain

The biosorption abilities of Rhodococcus opacus were studied for cadmium and zinc removal for liquid aqueous streams. The influence of pH, initial metal concentration and time removal were evaluated on the biosorption studies, in a batch scale basis. The Cd²⁺ and Zn²⁺ species uptake capacity by R. opacus has been also compared using Langmuir and Freundlich models. At pH 7.0 and 26 °C C, Cd²⁺ removal achieved a value around 60% from an initial concentration of 15 ppm. On the other hand, Zn removal achieved a value around 88% from an initial concentration of 5 ppm. Kinetics studies revealed that the biosorption process followed a pseudo-second order model for the two metal species (Cd²⁺ and Zn²⁺) and the kinetic constants were 3.90 and 3.37 g mg⁻¹ min⁻¹, for an initial concentration of 15 and 5 ppm for cadmium and zinc, respectively. The results showed that the R. opacus is a potential engineering biosorbent for environmental and extractive metallurgy sustainable applications.     

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.     

A simple spectrophotometric method for determination of gold (III) in aqueous samples

A simple method for rapid determination of trace Au in natural water was presented by using UV–vis spectrophotometry after reaction of gold (III) with 3,3′, 5,5′-tetramethylbenzidine hydrochloride (TMBH) in acidic solution. Under the optimum conditions, in a concentration range of 100–2000 μg L^?1 of Au (III) a good linear calibration graph was obtained (r = 0.9969, n = 7). The percent relative standard deviation (RSD) for determination of 1000 μg L^?1 Au was 10% (n = 3) and limit of detection based on a signal-to-noise ratio (S/N) of 3 (3S_bl) was 50 μg L^?1. The proposed method has been successfully applied to the determination of gold spiked and real aqueous samples.     

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.     

Biological treatment of wastewater produced during recycling of spent lithium primary battery

Wastewater produced during recycling of spent lithium primary battery was biologically treated with Acidithiobacillus ferrooxidans to decrease the pH and metal concentration. Since the wastewater contains high concentrations of Cr, Ni, and Li, the effects of these metals on the bacterial activity in a 9 K medium were also investigated. Samples of the medium with different metal concentrations were treated, and the oxidation ratio of Fe²⁺ ions was measured to examine the activity of bacteria. In the treatment of simulated wastewater, the presence of Cr and Ni ions with concentrations of 8000 g m^?3 and 13,000 g m^?3, respectively, did not inhibit the bacterial activity, whereas the oxidation ratio of Fe²⁺ ions was observed to be low in the medium when Li ion was present with a concentration at 5000 g m^?3. This observation suggested that at this concentration, Li ion suppressed the bacterial activity. In the case of treatment of real wastewater containing Cr, Ni, and Li, the oxidation ratio of Fe²⁺ to Fe³⁺ was observed to be low while the Fe concentration and pH decreased to 21,633 g m^?3 and 1.8, respectively. Thus, the wastewater produced during the recycling of spent lithium primary batteries can be effectively treated biologically for re-circulating in the recycling process.     

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.     

Copper leaching from waste electric cables by biohydrometallurgy

This study examines the leaching of copper from waste electric cables by chemical leaching and leaching catalysed by Acidithiobacillus ferrooxidans in terms of leaching kinetics and reagents consumption. Operational parameters such as the nature of the oxidant (Fe³⁺, O₂), the initial ferric iron concentration (0–10 g/L) and the temperature (21–50 °C) were identified to have an important influence on the degree of copper solubilisation. At optimal process conditions, copper extraction above 90% was achieved in both leaching systems, with a leaching duration of 1 day. The bacterial leaching system slightly outperformed the chemical one but the positive effect of regeneration of Fe³⁺ was limited. It appears that the Fe²⁺ bio-oxidation is not sufficiently optimised. Best results in terms of copper solubilisation kinetics were obtained for the abiotic test at 50 °C and for the biotic test at 35 °C. Moreover, the study showed that in same operating conditions, a lower acid consumption was recorded for the biotic test than for the abiotic test.     

Review of rate constants for thiosulphate leaching of gold from ores,concentrates and flat surfaces: Effect of host minerals and pH

A review of literature data for different types of sulphide concentrates and gold ores has been carried out to examine the impact of host minerals and pH upon gold leaching. Analysis of initial rate data over the first 30–60 min of gold leaching from sulphide concentrates or silicate ores over a range of ammonia, thiosulphate, and copper(II) concentrations, pH (9–10.5) and temperatures up to 70 °C shows the applicability of a shrinking sphere kinetic model with an apparent rate constant of the order k_ss = 10⁻⁶–10⁻³ s⁻¹. The dependence of apparent rate constant on pH and initial concentrations of copper(II) and thiosulphate is used to determine a rate constant k_Au(ρr)⁻¹ of the order 1.0 × 10⁻⁴–7.4 × 10⁻⁴ s⁻¹ for the leaching of gold over the temperature range 25–50 °C (ρ = molar density of gold, r = particle radius). These values are in reasonable agreement with rate constants based on electrochemical and chemical dissolution of flat gold surfaces: k_Au = 1.7 × 10⁻⁴–4.2 × 10⁻⁴ mol m⁻² s⁻¹ over the temperature range 25–30 °C. The discrepancies reflect differences in surface roughness, particle size and the effect of host minerals.     

Iron redox-equilibria and sulphide capacity of PGM melter-type slags

New measurements have been made on the ferric to ferrous ratio as well as the sulphide capacity for platinum group metals (PGM) melter-type slags. In South Africa, these slags are produced from the smelting of low-grade copper–nickel sulphide ores, Nell [Nell, J., 2004. Melting of platinum group metal concentrates in South Africa. The South African institute of Mining and Metallurgy 104 (7), 423–428]. The typical mass compositions are 5–10% Al₂O₃, 2–15% CaO, 5–30% FeO_x, 15–25% MgO and 40–60% SiO₂ with a molar basicity defined as (CaO + MgO)/SiO₂ of 0.6–1. The industrial furnaces operate at temperatures ranging from 1450 to 1600 °C under fairly reducing conditions (typically a pO₂ close to 10⁻⁸ atm at 1500 °C). The gas–slag equilibrium was studied by subjecting a synthetic slag to controlled atmospheres in a vertical tube-furnace using Ar–CO–CO₂ (–SO₂) gas mixtures. The ratio of ferric to ferrous was determined at 1450 °C for oxygen activities, defined as pCO₂/pCO, ranging from 0.11 to 1.75 by analysing the quenched slags using the standard titration and XRF techniques. The measured Fe³⁺/Fe²⁺ ratio increased from 0.029 to 0.110 with the increasing oxygen activity. Slight non-ideal iron redox behaviour was observed, as has been reported for low alumina and low iron-containing slags. The present results are in good agreement with the trends found in the literature for similar multi-component slag systems (mostly iron bath smelting slags). Sulphide capacity was measured at partial pressures of oxygen and sulphur of approximately 10⁻⁹ and 10⁻³ atm respectively, with total-iron contents of 8.2 and 15.6 wt%, and temperature ranging from 1450 to 1525 °C. The present sulphide capacity data ranged from 10^−4.43 to 10^−3.71. The expected increase in sulphide capacity with increasing temperature was observed, and at a given temperature, the sulphide capacity increased with an increase in iron oxide content.