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.     

Stripping conditions to prevent the accumulation of rare earth elements and iron on the organic phase in the solvent extraction circuit at Skorpion Zinc

The concentrations of the rare earth elements (particularly Y, Yb, Er and Sc) in the zinc-stripped organic phase streams in the solvent extraction process at Skorpion Zinc mine have increased gradually over the past four years. Iron is the only other impurity present in notable quantities in the organic phase after washing and scrubbing prior to zinc stripping. This project aimed to evaluate the effects that rare earth elements and iron in the organic phase have on the zinc solvent extraction process and to subsequently find appropriate stripping conditions for the removal of these elements from the zinc-stripped organic phase.Results obtained by performing laboratory scale batch tests indicated that the viscosity of the organic phase doubled and the phase disengagement time increased from 100 s to 700 s when the total rare earth elements and iron concentration in the organic phase was increased from 3100 mg/L to 6350 mg/L. The zinc loading capacity of the organic phase after two extractions furthermore decreased by a value between 1 g/L and 3 g/L, depending on the composition of the pregnant leach solution. The stripping of low concentrations of rare earth elements and iron from 40% di-2-ethylhexyl phosphoric acid (D2EHPA) diluted in kerosene was evaluated using two different stripping agents (H₂SO₄ and HCl) with concentrations between 1 M and 7 M, organic-to-aqueous ratios between 0.5 and 4, and temperatures between 30 °C and 50 °C. The highest stripping percentages were achieved at acid concentrations greater than 5 M, organic-to-aqueous ratios of less than 0.5 and high temperatures.     

Study on separation of heavy rare earth elements by solvent extraction with organophosphorus acids and amine reagents

The present work describes a study of the separation of rare earth elements (REE) from heavy REE concentrate through solvent extraction. Seven extractants were investigated: three organophosphorus acids (DEHPA, IONQUEST^®801 and CYANEX^®272), a mixture of DEHPA/TOPO (neutral ester) and three amines (ALAMINE^®336, ALIQUAT^®336 and PRIMENE^®JM-T). The organophosphorus extractants were investigated in hydrochloric and sulphuric media whereas the amines performance was assessed in a sulphuric medium. The variables investigated were: concentration of the extractant agent, aqueous phase acidity, aqueous/organic volumetric ratio, contact time, stripping agent concentration (hydrochloric acid solution) and the selective stripping step. In the extraction step, the best separation factors for the adjacent elements were obtained with DEHPA and IONQUEST 801. For 1.0 mol L⁻¹ DEHPA in an initial acidity of 0.3 mol L⁻¹ H⁺, the separation factor was 2.5 Tb/Dy, 2.1 Dy/Ho, 1.9 Ho/Er, 2.0 Ho/Y and 1.1 Y/Er; for 1 mol L⁻¹ IONQUEST 801 in 0.3 mol L⁻¹ of H⁺ it was 2.7 Tb/Dy, 2.4 Dy/Ho, 2.1 Ho/Er, 2.1 Ho/Y e 1.5 Y/Er. The study concluded that for the extractants investigated, IONQUEST 801 is the most indicated for the separation of heavy REE because it has lower affinity with the REE compared to the affinity of DEPHA/REE, which makes the strip of the REE from Ionquest 801 easier than from DEHPA. Moreover, the number of stages necessary for the stripping of the REE from IONQUEST 801 is much lower than that observed when DEPHA is employed.     

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.     

Separation of nickel from copper in ammoniacal/ammonium carbonate solution using ACORGA M5640 by selective stripping

Separation of nickel from copper in ammoniacal/ammonium carbonate solution using ACORGA M5640 by selective stripping was carried out. The influence of equilibration time, equilibrium pH and extractant concentration on the extraction of both the metals was studied. It was found that the copper extraction equilibrium was reached in a shorter time than the nickel extraction equilibrium. Nickel extraction decreases above an equilibrium pH of 9.0, while the extraction of copper remains unaffected by the changes in the equilibrium pH range of 7–10. Co-extraction, ammonia scrubbing and the selective stripping of copper and nickel were performed for a solution containing 3 g/l each of copper and nickel and 60 g/l ammonium carbonate. The extraction and the percentage stripping of copper and nickel were almost quantitative.     

Preparation of layered nickel aluminium double hydroxide from waste solution of nickel

The separation of nickel has been carried out from a waste solution containing 3.18 g/L Ni with other impurities such as Fe, Zn, Cu and As. Iron was removed by precipitation and Cu and Zn were removed by solvent extraction using LIX 622N and NaTOPS-99, respectively. After removal of all these impurities nickel was extracted by 1.5 M NaTOPS-99 in two counter-current stages at A:O ratio of 3:1 and the loaded organic was stripped with 30 g/L H₂SO₄ at phase ratio of unity. The strip solution of nickel was treated with Al₂(NO)₃ · 9H₂O for co-precipitation by increasing the pH of solution with 1 M NaOH up to 10. The Ni–Al layered double hydroxide was confirmed through XRD characterization.     

Solvent extraction separation and recovery of cobalt and nickel from sulphate medium using mixtures of TOPS 99 and TIBPS extractants

We report in this paper the solvent extraction separation of cobalt and nickel from synthetic sulphate solutions using TOPS 99 and TIBPS mixtures diluted in kerosene. The feed contains 1.061 g/L Co and 1.187 g/L Ni. Extraction experiments with synergistic mixture of extractants showed highest separation factor of 12,245 with 0.1 M TOPS 99 and 0.05 M TIBPS at pH 1.1. McCabe–Thiele plot for Co extraction with 0.1 M TOPS 99 and 0.05 M TIBPS extractants mixture indicated the necessity of three theoretical stages for >99% Co extraction at an aqueous to organic phase (A/O) ratio of 2. A three stage counter current extraction simulation test conducted at pH 1.1 with 0.1 M TOPS 99 and 0.05 M TIBPS mixture, confirmed Co extraction of 99.5% with Ni co-extraction of 0.02%. The results demonstrated that the addition of TIBPS–TOPS 99 acts as a synergist for Co extraction and antagonist for Ni.     

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.     

Use of quaternary ammonium salts to remove copper–cyanide complexes by solvent extraction

Cyanidation is one of the most common methods for the extraction of precious metals. In this process, effluents frequently contain relatively high concentrations of copper, which may react with cyanide to form cuprocyanide complexes adversely affecting the process.In this preliminary work, the use of solvent extraction to remove the copper–cyanide species from a synthetic solution similar to that of gold mill effluents was studied in order to permit the recycling of the solution into the process. For the extraction of these anions, the quaternary ammonium salts Quartamin TPR, Adogen 464 and Aliquat 336 were studied as extractants. The experimental results showed that for a synthetic solution of 710 mg/L copper and 1100 mg/L cyanide, it is possible to obtain a copper extraction of 99% when using 0.033 mol/L of the extractant Adogen 464 (organic/aqueous volume ratio (O/A) = 1) in the range of pH of 9–11. Up to 99% of the copper can be stripped from the organic solution after three contact times (5 min each) with 50 mL of sodium hydroxide 0.5 M (O/A = 1).     

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.     

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.     

Metal ions released from fluid inclusions of quartz associated with sulfides

The release of fluid inclusions has a strong potential for the unintentional activation of minerals during flotation. The present study aims to characterize fluid inclusions in natural quartz from a complex sulfide ore deposit. The results indicate that many fluid inclusions exist in the quartz. Under the experimental conditions of 2 g of quartz cleaned in 40 ml of pure deionized water under an inert atmosphere, the concentrations of Cu, Pb, Zn and Fe in aqueous solution reach concentrations of 1.92 × 10⁻⁷, 8.88 × 10⁻⁷, 8.31 × 10⁻⁷ and 90.33 × 10⁻⁷ mol/L, respectively. These values are significantly greater than those from the experimental non-oxidative dissolution of the quartz. In addition, the concentrations of metal ion released from fluid inclusions in the quartz sample at conditions approached “typical” industrial flotation environment are determined. The results indicate that the fluid inclusions of quartz represent the considerable sources of Cu, Pb, Zn and Fe in the aqueous solution. The present investigation provides a new understanding for the source of the unavoidable metal ions in the flotation pulp and may benefit understanding of the flotation theory.     

Preparation of immobilized sulfate reducing bacteria (SRB) granules for effective bioremediation of acid mine drainage and bacterial community analysis

Heavy metal-resistant immobilized sulfate-reducing bacteria (SRB) granules were prepared to treat acid mine drainage (AMD) containing high concentrations of multiple heavy metal ions using an up-flow anaerobic packed-bed bioreactor. The bioreactor demonstrated satisfactory performance at influent pH 2.8 and high concentrations of metals (Fe 463 mg/L, Mn 79 mg/L, Cu 76 mg/L, Cd 58 mg/L and Zn 118 mg/L). The effluent pH ranged from 7.8 to 8.3 and the removal efficiencies of Fe, Cu, Zn and Cd were over 99.9% except for Mn (42.1–99.3%). The bacterial community in the bioreactor was diverse and included fermentative bacteria and SRB (Desulfovibrio desulfiricans) involved in sulfate reduction. The co-existing anaerobic fermentative bacteria (Clostridia bacterium, etc.) with the ability to use lactate as electron donor could explain the differences between actual lactate consumption and what would be expected based solely on sulfate reduction.     

An environmentally-friendly technology of vanadium extraction from stone coal

A novel technology characterized by higher recovery of vanadium and which was environmentally-friendly was developed to recover vanadium from stone coal. Vanadium in stone coal could be leached by NaOH solution after roasting stone coal at 850 °C for 3 h. H₂SO₄, Mg(NO₃)₂ and ammonia were employed, respectively, in two steps to remove the impurities of Si and Al from the leach liquor. After extracting vanadium from the leach liquor with 10 vol% N₂₃₅, 20 vol% secondary octyl alcohol and 70 vol% sulfonated kerosene, 1.5 mol/L NaOH was used as a stripping agent to strip vanadium from extracting solution. Adding 80 g/L NH₄NO₃ to the stripping solution at 30–40 °C and pH 7.5, vanadium could be crystallized as ammonium metavanadate. Roasting ammonium metavanadate at 540 °C for 1 h, the purity of V₂O₅ met the standard specification. The total recovery of vanadium reached 67.39%, which was higher than the classical technology.     

Cobalt and zinc recovery from copper sulphate solution by solvent extraction

It has been demonstrated in earlier works that zinc as an impurity can be effectively removed from cobalt sulphate solutions (Zn/Co < 1) by solvent extraction with D2EHPA. Some process residues from copper plants contain both cobalt and zinc as valuable metals, which have to be separately extracted for their recovery. Leaching of such residues leads to solutions with higher Zn/Co ratios (Zn/Co > 10). Again, solvent extraction with D2EHPA has been successfully used to separate cobalt and zinc into their respective solutions, which could further be treated by appropriate techniques for the production of these metals.The method mainly consists of selective copper extraction with LIX 984, iron removal by precipitation with CaCO₃, simultaneous cobalt and zinc extraction with D2EHPA followed by their separation by selective stripping with sulphuric acid of different concentrations. The use of a specific cobalt extractant is not necessary. More than 95% copper has been recovered from the pregnant solution typically containing 1.0 g/l Co²⁺, 2.0 g/l Cu²⁺, 12.60 g/l Zn²⁺ and 8.4 g/l Fe³⁺. The cobalt and zinc recoveries were on an average of 90% each in their respective individual solutions.     

Biosorptive flotation of copper ions from dilute solution using BSA-coated bubbles

Copper adsorption was carried out using the novel material known as air-filled emulsion (AFE). AFE is a stable colloidal system containing microscopic protein-coated bubbles (<10 μm) dispersed through an aqueous solution, resulting in an increased specific surface area and contact time between extractant and metal ions. Bovine serum albumen (BSA) generated emulsion concentration had a significant impact on copper removal, with maximum metal uptake obtained at 2.5 g/l of BSA-coated bubbles. It was shown that copper sorption was rapid over the first 10 min, and equilibrium conditions were reached within 40 min. Separation of the copper-loaded microcells from the aqueous solution was also investigated. Micro-flotation was employed to remove the microbubbles by means of attachment to the surface of larger air bubbles. In absence of a cationic surfactant, approximately 0.5% copper recovery was obtained at pH ranging from 5 to 8 due to the lack of hydrophobic groups on the surface of Cu-loaded BSA emulsions. Due to the fine sizes of the emulsion bubbles (<10 μm) a cationic flocculant was used to induce coagulation of the bubbles leading to easier phase separation. A combination of collector and flocculant at a concentration of 3 × 10⁻⁴ M and 0.025 g/l, respectively, led to an increase in copper recovery to nearly 35% at pH 7.     

Solvent extraction of tetravalent hafnium from acidic chloride solutions using 2-ethyl hexyl phosphonic acid mono-2-ethyl hexyl ester (PC-88A)

Solvent extraction of Hf(IV) from acidic chloride solutions has been carried out with PC-88A as an extractant. Increase of acid concentration decreases the percentage extraction of metal indicating the ion exchange type mechanism. The plot of logD vs log[extractant], M is linear with slope 1.8 indicating the association of two moles of extractant with the extracted metal species. Plot of logD vs log[H⁺] gave a straight line with a negative slope of ∼2 indicating the exchange of two moles of hydrogen ions for every mole of Hf(IV). The effect of Cl⁻ ion concentration at constant concentration of [H⁺] did not show any change in D values. Addition of sodium salts enhanced the percentage extraction of metal and follows the order NaSCN > NaCl > NaNO₃ ∼ Na₂SO₄. Stripping of metal from the loaded organic (LO) with different acids indicated sulphuric acid as the best stripping agent. Regeneration and recycling capacity of PC-88A, temperature, extraction behavior of associated elements was studied.     

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%.