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.     

Use of mesalime and electric arc furnace (EAF) dust as neutralising agents in biooxidation and their effects on gold recovery in subsequent cyanidation

The cost of lime/limestone for neutralisation is the second largest operating cost in bioleaching. Therefore, these studies have been conducted with the aim to investigate the possibilities for use of by-products such as mesalime and electric arc furnace (EAF) dust for neutralisation during biooxidation of a refractory gold concentrate. Experiments were carried out using a retention time of 57 h in a one-stage reactor and the influence of two industrial by-products on the biooxidation performance was evaluated. The neutralising capacity of EAF dust was lower, while the mesalime was similar to the Ca(OH)₂ reference. The arsenopyrite oxidation in experiments ranged from 85% to 90%, whereas the pyrite oxidation was 63–74%. In subsequent cyanidation, final gold recoveries of 90% were achieved in bioresidues from mesalime and Ca(OH)₂, while the EAF dust bioresidue had a recovery of 85%. A comparatively high elemental sulphur content in EAF dust probably encapsulates part of the gold, which explains the lower recovery for the EAF dust bioresidue despite a longer residence time. Cyanide consumption was relatively high and ranged from 8.1 to 9.2 kg/ton feed after 24 h of cyanidation. Overall, the by-products tested here have proved to be feasible options as neutralising agents in bioleaching operations.     

Complex treatment of mining and metallurgical wastes for recovery of base metals

A complex process for the recovery of copper and zinc from mining and metallurgical wastes has been investigated and proposed. It includes sulfuric acid leaching of old pyrite flotation tailings to produce ferric containing leach solution; followed by ferric leaching of copper converter slag flotation tailings with the leach solution. A sample of old pyrite flotation tailings from the concentrator containing 0.36% of copper and 0.23% of zinc was leached with 10% sulfuric acid in the column. Recovery of copper and zinc reached 47.1% and 47.2%, respectively. The pregnant leach solutions contained 15.9 g/L of ferric iron. The subsequent ferric leaching of copper converter slag flotation tailings containing 0.53% copper and 2.77% zinc with the pregnant leach solution was conducted. The effects of various process parameters on the leaching dynamics of metals under batch conditions were investigated. Under the best conditions (temperature 70 °C, pulp density 30%, ferric iron concentration 15.9 g/L, initial pH of the pulp 0) the recovery of copper and zinc reached 79.6% and 43.7%, respectively. It was concluded that acid leaching of base metals from old pyrite flotation tailings with pregnant leach solution for the ferric leaching of copper converter slag flotation tailings is a prospective and promising technique for the complex treatment of mining and metallurgical wastes.     

Amenability for processing of oolitic iron ore concentrate for phosphorus removal

Beneficiation routes aimed at dephosphorisation of oolitic gravity magnetic concentrate and involving a combination of roasting, re-grinding, magnetic separation and water and acid leaching are investigated. Roasting was carried out at 900 °C for 1 h without or with lime or sodium hydroxide as roasting additives. When additives were used, cement phases of Si–Al–Na–Ca–O type were detected as well as the mineral giuseppettite. During the thermal process sodium silicate is liquefied and the newly formed phases coat the oolites and penetrate inside the cracks. Energy Dispersive Spectroscopy analysis has indicated that the zone surrounding the oolites consists of Na, Al and Si phases with part of phosphorus being captured there. As a result of the alkaline roasting, goethite is partly transformed to magnetite and this reduction is reinforced with an increase in sodium hydroxide dosage. Investigation of redistribution of phosphorous shows that it could be only partly separated if leaching is not accompanied by re-grinding and physical separation. The recommended dosage of the reductive agent for the final flowsheet is 8 mass% ratio to concentrate. Grinding to a mean size of 0.040 mm, with water and acid leaching and double magnetic separation creates conditions to obtain a high-quality iron concentrate with 65.97% Fe and recovery of 92.43%, with simultaneous decrease in the phosphorus content from 0.71% to 0.05%.     

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.     

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%.     

Acid mine drainage treatment through a two-step neutralization ferrite-formation process in northern Japan: Physical and chemical characterization of the sludge

The acid mine drainage of a closed iron sulphide mine in northern Japan was treated on-site using a continuous flow bench scale plant. In the bench scale plant of the two-step neutralization ferrite-formation process, magnesium oxide or calcium carbonate was used during the first neutralization step to raise the pH to around 4.8 to produce aluminium hydroxide sludge. In the second neutralization step sodium hydroxide was applied to reach a pH of 8.5 to produce ferrite sludge. Initial settling rates of the produced aluminium and ferrite sludge were from 1.2 to 4.8 times higher than the initial settling rates of the sludge produced by the treatment plant currently operated at the mine. The suspended solids (SS) concentration in the sludge ranged from 1.0 to 11.8 times higher than the SS concentration in the sludge produced by the current facility. The sludge volume index (SVI) of aluminium sludge was 11 and 36 mL/g when produced with magnesium oxide or calcium carbonate, respectively. The SVI of ferrite sludge was 4 mL/g regardless of the type of neutralizer used in the first neutralization, while the same parameter (SVI) of the sludge produced by the current facility is 70 mL/g, which indicates that the two-step neutralization ferrite-formation process generates sludge with much higher density. In addition, the process effectively reduced the concentration of toxic heavy metals from above 800 ppb, 13 ppm, and 15 ppm to as low as 1.4 ppb, 0.02 ppm and 0.2 ppm for arsenic, copper, and zinc, respectively.     

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.     

A novel method to recover zinc and iron from zinc leaching residue

A novel method to recover zinc and iron from zinc leaching residue (ZLR) by the combination of reduction roasting, acid leaching and magnetic separation was proposed. Zinc ferrite in the ZLR was selectively transformed to ZnO and Fe₃O₄ under CO, CO₂ and Ar atmosphere. Subsequently, acid leaching was carried out to dissolve zinc from reduced ZLR while iron was left in the residue and recovered by magnetic separation. The mineralogical changes of ZLR during the processes were characterized by XRF, TG, XRD, SEM–EDS and VSM. The effects of roasting and leaching conditions were investigated with the optimum conditions obtained as follows: roasted at 750 °C for 90 min with 8% CO and CO/CO + CO₂ ratio at 30%; leached at 35 °C for 60 min with 90 g/l sulfuric acid and liquid to solid ratio at 10:1. The iron was recovered by magnetic separation with magnetic intensity at 1160 G for 20 min. Under the optimum operation, 61.38% of zinc was recovered and 80.9% of iron recovery was achieved. This novel method not only realized the simultaneous recovery of zinc and iron but also solved the environmental problem caused by the storage of massive ZLR.     

Atmospheric acid leaching of siliceous goethitic Ni laterite ore: Effect of solid loading and temperature

In this study, atmospheric acid leaching behaviour of siliceous goethitic nickel (Ni) laterite ore is investigated. Specifically, the effect of −200 μm feed solid loading (30 vs. 45 wt.%) and temperature (70 vs. 90 °C) on leach kinetics, acid consumption capacity and Ni and cobalt (Co) extraction was studied under isothermal, batch (4 h) leaching conditions at pH 1. Incongruent leaching was observed for constituent elements reflecting slow but steady release of value (Ni and Co) and some of gangue metals such as Fe, Mg and Al accompanied by faster and sharp release of Na and Si. Higher temperature and lower pulp solid loading, both led to a 40–50% increase in overall Ni and/or Co extraction and higher acid consumption. At 70 °C and 45 wt.% solid loading, Ni/Co extraction after 4 h was the lowest (∼14/16%) whilst the highest extraction (∼67/56%) was observed at 90 °C and 30 wt.% solid loading. Temperature appeared to have dramatic influence on Ni/Co and other impurity metals’ extractions revealing the chemical reaction controlled nature of the leaching. Higher solid loading and longer leaching time also both slowed down the leach kinetics. A two-stage chemical reactions-controlled leaching mechanism involving a faster initial leaching kinetics followed by a slower leaching at lower rate constants and higher activation energies was established for release of Ni, Co, Fe and Mg. The mechanism reflects the fast leaching of reactive host mineral phases (e.g., clays and Mg–silicates) during first 30 min followed by slow leaching of more refractory mineral phases (e.g., goethite and quartz) during the rest of leaching period. The findings provide a greater understanding for enhanced atmospheric acid leaching process of siliceous goethitic laterite ores.     

Study of silver leaching with the thiosulfate–nitrite–copper alternative system: Effect of thiosulfate concentration and leaching temperature

Thiosulfate system is considered an interesting alternative leaching process for precious metals. Nevertheless, most of the literature published on these conventional thiosulfate leaching solutions has been focused on the use of ammonia and copper to generate the cupric tetraamine complex, which acts as a catalytic oxidant for silver. However, ammonia toxicity is also a detrimental issue in terms of the process sustainability. For that reason, thiosulfate–nitrite–copper solutions were studied as an alternative less toxic system for silver leaching.In this work, the effect of the thiosulfate concentration (0.07 M, 0.1 M and 0.15 M) and temperature (room temperature, 30, 35, 40 and 45 °C) on the metallic silver leaching kinetics is presented for the S₂O₃–NO₂–Cu system. The results show that the thiosulfate concentration plays an important role in the S₂O₃–NO₂–Cu–Ag system since it controls the silver leaching kinetics. On the other hand, an increase in temperature favors the silver recovery.Finally, the SEM–EDS analysis, the X-ray mapping and the X-ray diffractograms show that the solid silver particles are coated by a Cu, S and O layer for the 0.07 M and 0.1 M thiosulfate experiments, which is consistent with the formation of antlerite (Cu₃(SO₄)(OH)₄); while the 0.15 M thiosulfate scenario produced a layer composed only of Cu and S, revealing the formation of stromeyerite (CuAgS). The UV–Visible technique confirmed the in-situ generation of copper–ammonia complexes for the 0.07 M leaching condition; however, these complexes are not formed at the 0.15 M condition.     

The removal of zinc from liquid streams by electroflotation

The removal of heavy metals from dilute aqueous solutions (in the range of 10⁻⁷–10⁻⁴ mol dm⁻³) is often not acceptable using classical methods, which do not achieve levels in accordance with environmental quality standards. Electroflotation has certain desirable characteristics, compared to dissolved and dispersed air flotation, particularly in regard to the small bubble size distribution of the process. The aim of this work was to develop an electroflotation (EF)/electrocoagulation (EC) cell to study this combined process and the influence of some relevant parameters/variables, such as collector concentration, tension and current density variation, on the removal of zinc from synthetic solutions containing 20 mg l⁻¹ of the metal. A platinum gore (5 mm) anode and stainless steel mesh cathode were used in the electroflotation cell. The work showed that it was possible to remove zinc by electroflotation, 96% removal being achieved using sodium dodecyl sulfate (SDS) as collector in the stoichiometric ratio 1:3, current density of around 8 mA/cm² and an inlet pH of about 7.0.     

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).     

Process development for the recovery of europium and yttrium from computer monitor screens

This work describes the development of a process for the recovery of Eu and Y from cathode ray tubes (CRTs) of discarded computer monitors with the proposition of a flow sheet for the metals dissolution. Amongst other elements, europium and yttrium are presented in the CRTs in quantities – 0.73 w/w% of Eu and 13.4 w/w% of Y – that make their recovery worthwhile. The process developed is comprised of the sample acid digestion with concentrated sulphuric acid followed by water dynamic leaching at room temperature. In the CRTs, yttrium is present as oxysulphide (Y₂O₂S) and europium is an associated element – Y₂O₂S:Eu³⁺ (red phosphor compound). During the sulphuric acid digestion, oxysulphide is converted into a trivalent Eu and Y sulphate, in solid form, with the liberation of H₂S. In the second step, metals are leached from the solid produced in the acid digestion step by dynamic leaching with water. This study indicates that a proportion of 1250 g of acid per kg of the sample is enough to convert Eu and Y oxysulphide into sulphate. After 15 min of acid digestion and 1.0 h of water leaching, a pregnant sulphuric liquor containing 17 g L⁻¹ Y and 0.71 g L⁻¹ Eu was obtained indicating yield recovery of Eu and Y of 96% and 98%, respectively. Both steps (acid digestion and water leaching) may be performed at room temperature.     

Metals recovery from galvanic sludge by sulfate roasting and thiosulfate leaching

The purpose of this work is the selective recovery of Au, Ag, Cu, and Zn from two types of galvanic sludge using a mixed process of sulfate roasting and sodium thiosulfate leaching. In the experiments, the sludge was mixed with a sulfate promoter (sulfur, iron sulfate, or pyrite) and treated by pyrometallurgical processes at temperatures up to 750 °C. At this stage, this agent is thermally oxidized, turning the furnace atmosphere into a reducing one and the metallic oxides into water-soluble sulfates. Afterward, the sulfates can be treated by leaching with water for recovery of Ag, Cu, and Zn. The gold does not form sulfates in this reaction and was recovered through a second leaching stage using sodium thiosulfate, an effective reagent and less harmful to the environment than cyanide. Different parameters such as the sulfate promoter that achieves the highest recovery of metals, the proportion of galvanic sludge to sulfating agent, the temperature, the heating time in the oven, and the leaching time were evaluated. Additionally, a comparison of gold recovery using cyanide versus sodium thiosulfate was performed. The configuration that showed the best metal recovery included a 1:0.4 ratio of sludge to sulfur, an oven temperature of 550 °C, a roasting time of 90 min, and a water leaching time of 15 min. Using these parameters, recovery rates of 80% of the silver, 63% of the copper, and 73% of the Zn were obtained. The sodium thiosulfate leaching resulted in a recovery of 77% of the Au, close to the values obtained using cyanide.     

Hydrothermal modification to improve the floatability of ZnS crystals

Crystalline structure and surface properties significantly affect the floatability of metal sulphides. In this study, a novel methodology to modify zinc sulphide (ZnS) crystals was proposed to improve the floatability of the crystals. Initially, ZnS crystals, synthesised from zinc hydroxide (Zn(OH)₂) and sulphur (S) under hydrothermal conditions, were used to assess the floatability. X-ray diffraction (XRD) and transmission electron microscopy (TEM) were employed to analyse the crystalline structure and surface properties of the sulphides. Conventional flotation tests were performed to evaluate the floatability. The effects of mineraliser (KOH) concentration, precursor (Zn(OH)₂) concentration, hydrothermal temperature and holding time on the floatability of the ZnS crystals were investigated. The optimal flotation recovery of ZnS (82.53%) was obtained with a KOH concentration of 5 mol/L, a Zn(OH)₂ concentration of 10%, a holding time of 4 h and a hydrothermal temperature of 260 °C. Then, sludge containing fine and amorphous zinc compounds, which was generated during the disposal of metallurgical waste water, was employed to test the recovery of valuable metals using modified hydrothermal sulphidation. The results show that the recovery of Zn in the sludge can reach 66.3% under the optimal conditions.     

A study of atmospheric acid leaching of a South African nickel laterite

Nickel and cobalt acid leaching from a low-grade South African saprolitic laterite using sulphuric acid was studied. Ore characterisation was performed by XRD and XRF. Batch agitation leaching tests were conducted at atmospheric pressure investigating main parameters: particle size and percent solids at 25 °C and 90 °C. Ore characterisation showed that the ore is a saprolitic laterite with nickel present in lizardite. Leaching tests showed that nickel and cobalt could be leached from the ore at atmospheric pressure. Nickel was found to be more leachable from the coarser −106 + 75 μm fraction, with 98% Ni being extracted at 90 °C after 480 min. Cobalt was not favoured by variation in particle size and increased percent solids. Increasing ore percent solids improved nickel extraction at 25 °C however at 90 °C extraction decreased due to a diffusion layer build-up as a result of amorphous colloidal silica. The co-dissolution of magnesium and iron was elucidated. Nickel leaching data at increased temperature and percent solids fit the shrinking core model equation, k_dt = 1−2/3x − (1 − x)^2/3 showing that nickel leaching reaction was diffusion controlled under the set conditions.     

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.     

Chloride assisted leaching of chalcocite by oxygenated sulphuric acid via Cu(II)–OH–Cl

The beneficial effect of the addition of sodium chloride upon the leaching kinetics of complex iron–nickel–copper sulphides at elevated temperatures and oxygen pressures has been widely reported since the late 1970s, but the role of chloride is still being investigated or debated. Previous researchers have considered chloride as: (i) a complexing agent for cuprous ions; (ii) a surfactant that disperses the molten sulphur and thus removes passivation of the mineral surface by elemental sulphur during pressure leaching; and (iii) a reagent which increases the surface area and the porosity of the insoluble product layer on the surface. A proper understanding of the role of chloride based on the leaching of individual sulphides of known composition in the absence of host minerals at low pulp densities would be useful for the development of chloride assisted sulphate leaching processes for complex sulphide ores, concentrates, and mattes. In the present study evidence for the formation of basic salts of Cu(II) and Fe(III) during leaching are presented. The published rate data are analysed for the leaching of copper from mono-sized chalcocite particles in oxygenated sulphuric acid solutions maintained at 85 °C, a temperature lower than the melting point of sulphur. The initial leaching follows a shrinking particle (sphere) model, and the apparent rate constants are first order with respect to the concentration of dissolved oxygen and chloride. The intrinsic rate constant for the surface reaction (0.2 m s⁻¹) is two orders of magnitude larger than the calculated mass transfer coefficient of oxygen (3 × 10⁻³ m s⁻¹). The proposed reaction mechanism considers the formation of an interim Cu(II)(OH)Cl⁰ species which facilitates the leaching process.     

Research on the recycling of valuable metals in spent Al2O3-based catalyst

A new technology was developed to recover multiple valuable elements in the spent Al₂O₃-based catalyst by X-ray phase analysis and exploratory experiments. The experiment results showed: In the condition of roasting temperature of 750 °C and roasting time of 30 min, mol ratio of Na₂O: Al₂O₃ 1.2, the leaching rate of alumina, vanadium and molybdenum in the spent catalyst is 97.2%, 95.8% and 98.9%, respectively. Vanadium and molybdenum in sodium aluminate solution can be recovered by barium hydroxide and barium aluminate, the precipitation rate of vanadium and molybdenum is 94.8% and 92.6%. Al(OH)₃ is prepared from sodium aluminate solution with carbonation decomposition process, and the purity of Al₂O₃ is 99.9% after calcinations, the recovery of alumina can reach 90.6% in the whole process. The Ni–Co concentrate was leached by sulfuric acid, a nickel recovery of 98.2% and over 98.5% cobalt recovery was obtained respectively under the experimental condition of 30% (w/w) H₂SO₄, 80 °C, reaction time 4 h, liquid:solid ratio (8:1) by weight, stirring rate of 800 rpm.