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.     

Kinetics of saprolitic laterite leaching by sulphuric acid at atmospheric pressure

Mineralogical analyses of the saprolitic laterite material have been characterized by X-ray fluorescence (XRF), X-ray diffraction (XRD), Fourier transform infrared (FTIR) spectroscopy, thermal analysis, scanning electron microscope (SEM) and energy dispersive X-ray analysis (EDAX). Results showed that the saprolitic laterite material consists mainly of nickel-substituted lizardite showing the pebble-like morphology and traces of magnetite and phlogopite. Leaching results showed that as much as 84.8% nickel could be leached under the experimental conditions of 10% (v/v) H₂SO₄, 90 °C reaction temperature, leached within 5 min, particle size d₅₀ = 25 μm, stirring at 500 rpm and liquid to solid ratio 3:1. The kinetics of nickel and magnesium leaching from the saprolitic laterite material have been investigated in a mechanically stirred reactor and the activation energies were determined to be 53.9 kJ mol^?1 for nickel and 59.4 kJ mol^?1 for magnesium respectively, which are characteristic for a chemical reaction controlled process. The similarity of the activation energies of nickel and magnesium leaching from the saprolitic laterite material by sulphuric acid means that nickel in lizardite is loosely bound within the octahedral layer and almost all of the nickel could be leached simultaneously with magnesium but without complete decomposition of the silicate structure.     

Leaching of manganese ores using sawdust as a reductant

In this study sawdust was used as reductant for sulphuric acid leaching of manganese ore. Effects of pulp density, amount of acid, temperature, particle size of ore and amount of sawdust were studied. Manganese extraction of ∼98% was achieved under the conditions: leaching time 8 h, 5% H₂SO₄ (v/v), 10% pulp density, 90 °C and 5% sawdust (w/w), i.e. 0.5 g/g ore. Other Mn containing materials like low grade manganese ore, manganese nodule and Mn-nodule leach residues were tested and all these materials responded well giving more than 98% Mn extraction.     

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.     

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.     

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.     

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.     

Innovative technology for processing saprolitic laterite ores by hydrochloric acid atmospheric pressure leaching

An innovative technology for processing saprolitic laterite ores from the Philippines by hydrochloric acid atmospheric leaching and spray hydrolysis is proposed. The factors that affect the hydrochloric acid atmospheric leaching of the laterite ores and spray hydrolysis of the atmospheric acid leach solution are investigated. Experimental results show that the leaching of Ni, Fe, and Mg is 98.9 wt%, 97.8 wt%, and 80.9 wt%, respectively, under optimal acid leaching conditions. The hydrolysis of Ni and Fe by the atmospheric acid leach solution approaches 100 wt% at the temperature range of 450–500 °C. Characterization results show that a serpentine mineral, nominally Mg₃Si₂O₅(OH)₄, is the major component and goethite, FeO(OH), is the minor one in the laterite ores. Treatment by hydrochloric acid atmospheric leaching breaks the mineral lattices of the laterite ores and makes amorphous silica the primary product in the atmospheric acid leach residue. The grade of Ni in the hydrolyzate increases to 4.55%. The hydrolyzate with high Ni content can be utilized for ferro-nickel production.     

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.     

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.     

Leaching behaviour of natural and heat-treated brannerite-containing uranium ores in sulphate solutions with iron(III)

Uranium leaching tests were conducted on two naturally occurring, highly metamict brannerite ores from the Crockers Well and Roxby Downs deposits, South Australia. The ores were leached over a range of temperatures and Fe^(III) and H₂SO₄ concentrations. As well, samples of the ores were calcined at 1200 °C in air to investigate the effect of thermally induced recrystallisation on uranium dissolution. For the unheated samples, a maximum of ∼80% U dissolution was obtained using an Fe^(III) concentration of 12 g/L, an acid concentration of 150 g/L H₂SO₄ and a temperature of 95 °C. The heat treated samples performed poorly under identical conditions, with maximum uranium dissolution of <10% recorded. High uranium dissolution from natural brannerite can be achieved providing; (i) acid strength, oxidant strength and temperatures are maintained at elevated levels (compared to those traditionally used for uraninite leaching), and, (ii) the brannerite has not undergone any significant recrystallisation (e.g. through metamorphism).     

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.     

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.     

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.     

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.     

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.     

Impact of organic carbon on the leachability of vanadium,manganese, iron and molybdenum from shale residues

From 1942 to the 1966, oil was produced by pyrolysis of shale, in Kvarntorp, Sweden. This generated some 40 million m³ of metal rich pyrolyzed shale and discarded fines that were piled on site with its original metal content almost intact. The present study focuses on the leaching of vanadium, manganese, iron and molybdenum from fines after addition of wood chips and steel slag, in outdoor 1 m³ reactor systems at low liquid to solid ratio, in order to evaluate the potential environmental impact and recovery of the elements from the leachates. Seasonal variations were observed, with increased leaching during peak summer. For vanadium and molybdenum, high addition of wood chips decreased the leaching, probably due to adsorption. Manganese showed the opposite behavior while leaching of iron was almost independent of the amount of wood chips. Depending on the systems, up to 2200 μg L⁻¹ vanadium, 90 μg L⁻¹ molybdenum, 25 mg L⁻¹ manganese and 500 mg L⁻¹ iron was found in the aqueous phase. Applied to the 40 million m³ pile, the annual leaching of those elements may reach 14 ton, 0.6 ton, 200 ton and 2400 ton, respectively.     

Agglomeration and column leaching behaviour of goethitic and saprolitic nickel laterite ores

Processability of complex, low-grade nickel (Ni) laterite ores via heap leaching is very limited due to some intractable geotechnical and hydrological challenges such as poor heap porosity/permeability and structural stability. This work presents some investigations on laboratory batch drum agglomeration and continuous column leaching behaviour of saprolitic (SAP) and goethitic (G) Ni laterite ores as part of the quest for an effective ore pre-treatment process for enhanced heap leaching. As a focus, the effect of ore mineralogy/chemistry on the agglomeration and column leaching behaviour of −2 mm (crushed from −15 mm run-of-mine) G and SAP Ni laterite ores was examined. To produce ∼5–40 mm agglomerates in <15 min, the SAP ore required a higher H₂SO₄ (30 wt.%) binder dosage compared with the G ore, although both ores displayed substantially similar, coalescence-controlled agglomeration mechanism. The resulting G agglomerates were more robust than the SAP ones based upon their compressive strength and acidic solution soak test measurements. However, over 100 days of continuous column leaching, the structural stability of the SAP agglomerate bed was slightly greater than that of G agglomerates, reflecting a lesser slump of the former. The pregnant leach solution analysis revealed greater Ni/Co extraction rates from the SAP than the G agglomerates. Whilst the total mass of acid consumed per ton dry ore processed was greater for the SAP ore, the total kg acid per kg Ni extracted was markedly lower. Incongruent leaching of gangue minerals’ constituent elements (e.g., Fe, Mn, Mg, Al and Si) occurred and contributed significantly to the overall acid consumption. The findings show the relevance of agglomeration and column leaching tests for providing useful information for plant designing and optimization of Ni laterite heap leaching operations.     

The effect of seawater based media on copper dissolution from low-grade copper ore

Due to the scarcity of water in the north of Chile, there is interest in small-scale mining using seawater to leach the minerals. This situation has led to this research aimed to determine the effect of different process variables on the extraction of copper based on the ore type of these small-scale mining activities. The extraction of copper from finely ground (<150 μm) low-grade mixed ore (0.36% Cu) was studied in different acidic media (H₂SO₄ and HCl). The effects of water quality (tap water, seawater and synthetic process water) and lixiviants on copper leaching were investigated at three temperatures (25 °C, 35 °C and 45 °C). Synthetic process water was prepared by adding Na⁺, Cl⁻, Cu²⁺ and SO₄²⁻ to seawater. Copper extractions between 70% and 80% were achieved in 48 h using seawater, similar to the extractions obtained using tap water. Rapid copper dissolution occurred when synthetic process water was used, from 90% to 97%. This marked increase was related to the addition of Cu²⁺, which promoted the formation of CuCl⁺. Seawater was comparable to freshwater in terms of leaching kinetics and yield potential by raising the chloride concentration and increasing the formation of copper chloride ions. The findings of this study also expanded our understanding of the consequences of substituting seawater for freshwater at industrial leaching operations.