Leaching of limonitic laterite ore by acidic thiosulfate solution

Cobalt is usually recovered as a by-product of copper and nickel processing, and only a small amount of cobalt is derived from laterites although a vast majority of cobalt resources in them. The exploitation of limonitic laterite containing high content of cobalt is becoming increasingly important. The mineralogy of a limonitic laterite ore was characterized by environmental scanning electron microscope (ESEM) and X-ray diffraction (XRD) in this paper. The results show that nickel occurs in goethite mainly, while cobalt is predominantly associated with manganiferous minerals. Thiosulfate is found to be able to selectively leach cobalt from limonitic laterite in the presence of sulfuric acid, and 91% Co, 22% Ni, 10% Fe are leached from an ore containing 0.13% Co, 1.03% Ni within the first 5 min at 90 °C under the conditions of 10 g/L sodium thiosulfate, 8% (w/w) sulfuric acid and 10:1 L/S ratio. The leaching kinetics of Mn and Co by acidic sodium thiosulfate solution can be characterized by the Avrami equation. In acidic solution, thiosulfate readily decomposes into sulfur and sulfur dioxide as intermediary reagents to reduce pyrolusite (MnO₂) and goethite (FeOOH); therefore, nickel and cobalt associated with goethite and pyrolusite respectively are extracted due to reduction dissolution. Furthermore, cobalt is selectively leached over iron and nickel because pyrolusite is preferentially reduced by acidic thiosulfate rather than goethite. The novel process may give an alternative method to selectively recover cobalt as the primary product from limonitic laterites at atmospheric pressure.     

Biomining in reverse gear: Using bacteria to extract metals from oxidised ores

Biomining, as traditionally practised, uses aerobic, acidophilic microorganisms to accelerate the oxidative dissolution of sulfide minerals present in ores and concentrates, thereby either causing target metals to be solubilised (e.g. copper) or made accessible to chemical extraction (e.g. gold). Many acidophiles are also able to catalyse the dissimilatory reduction of ferric iron in anoxic or oxygen-depleted environments, and can accelerate the reductive dissolution of ferric iron minerals, such as goethite, under such conditions. Recent work has demonstrated how this approach can be used to extract metals (nickel, copper, cobalt and manganese) from oxidised ores, such as laterites deposits, at low (∼30 °C) temperatures. Reductive mineral dissolution has been trialled successfully with a variety of ores, pointing to a generic application of this approach.     

Activation pretreatment of limonitic laterite ores by alkali-roasting method using sodium carbonate

Activation pretreatment of Cr-containing limonitic laterite ores by Na₂CO₃ roasting to remove Cr and Al, as well as its effect on Ni and Co extraction in the subsequent pressure acid leaching process were investigated. X-ray diffraction (XRD), thermogravimetric (TG), and scanning electron microscopy/X-ray energy dispersive spectroscopy (SEM/XEDS) techniques were used to characterize the laterite ores and the water leaching residues of alkali roasting and found that goethite is the major Ni-bearing mineral and chromite the minor one. Alkali-roasting pretreatment breaks the mineral lattices of the laterite, exposing their Ni and Co, which leads to higher extraction of these two metals under milder operation conditions in the subsequent pressure acid leaching process. Experimental results showed that the leaching of Cr and Al were up to 99 wt% and 80 wt%, respectively, under optimal alkali roasting and water leaching conditions. Compared with the direct pressure acid leaching of the raw laterite ores, leaching of Ni and Co increased from 79.96 wt% to 97.52 wt% and 70.02 wt% to 95.33 wt%, respectively, after alkali-roasting activation pretreatment was performed. Meanwhile, the grade of acid leaching iron residues increased from 55.31 wt% to 62.92 wt%, and these residues with low Cr content could be more suitable as the raw materials for iron-making.     

Thermodynamic analysis of the carbothermic reduction roasting of a nickeliferous limonitic laterite ore

As the sulfide ore deposits become less economically viable as a source of nickel, increasing attention is being paid to the nickeliferous laterite ores. However, in contrast to the sulfide ores, these oxide ores cannot be as easily concentrated by current technologies. Consequently, considerable research effort is being directed at developing new techniques for beneficiating the nickeliferous laterites. The pyrometallurgical production of a high-grade ferronickel alloy using a low cost carbonaceous reductant at relatively low temperatures is particularly attractive. In the current research, a thermodynamic model has been developed to aid in the understanding of the carbothermic reduction roasting process as a potential upgrading method for the nickeliferous limonitic laterite ores. The effects of process parameters such as temperature and reductant to ore ratio on the grade of the ferronickel alloy produced and the nickel recovery in the alloy have been studied. The thermodynamic predictions are shown to be in general agreement with the experimental results currently available in the literature.     

Rheological behaviour of lateritic smectite ore slurries

The characterisation and rheology of several nickel laterite smectite ores and pure minerals are compared to assess the effect of mineralogy and particle size on the viscosity of high pulp density slurries. A vane viscometer was used to determine the “optimum pulp density” (OPD) that gave a yield stress of 100 Pa which is considered to be optimal for pumping slurries into autoclaves in the HPAL process. In general, slurries containing finer particles were more viscous and smectite slurries exhibited poor rheological behaviour as compared to slurries of goethite < kaolin < talc < hematite < maghemite < magnesite. Blending the smectite ores with a fraction of the pure minerals improved the rheological behaviour of the pulp and can increase the optimum pulp density of the smectite blend by over 5% w/w.When the physical properties of the smectite ore and slurry were examined, a very good linear correlation was obtained between the optimum pulp density and the settling density which provides a simple measure of predicting rheological behaviour of slurries. The variation in the viscosity of the nickel laterite ores depends largely on their mineralogy and particle size distribution. The mean particle size and P₈₀ values of various smectite ores containing the same mineral phases were also found to have a reasonably good linear correlation with OPD in saline water, but the correlation of ore surface area with OPD was a poorer fit.     

Dissolution behaviour of a Turkish lateritic nickel ore

The atmospheric pressure sulphuric acid leaching characteristics of Adatepe (Eski?ehir, Turkey) laterite ore that has recently been put into operation was investigated. The effects of sulphuric acid concentration (5-95%), temperature (20-95 °C) and time (30-240 min) on leaching were determined by nickel, iron and arsenic analyses. The amounts of Ni, Fe and As in solution were observed to increase with increase of temperature from 20 °C to 70 °C for sulphuric acid concentrations between 5% and 95%. Further increase of temperature to 95 °C showed that the dissolution of Ni, Fe and As were increased until 60% sulphuric acid concentration and over 60% a decrease in the dissolution percentages was observed due to the probable formation of nickel and silicon containing ferric sulphate type compounds that cause nickel loss from the leach solution. Experimental results showed that maximum nickel dissolution of 99.2% at 95 °C could be reached in 120 min of leaching time for a sulphuric acid concentration of 60%. The congruency of Ni dissolution with respect to Fe was found to be congruent over about 25% Ni and 15% Fe dissolution values. XRD analyses on the residues obtained after leaching showed that it was not required to dissolve all goethite phase to reach maximum dissolution of nickel contained in the sample. An activation energy of 30.36 kJ/mole was determined for Ni dissolution showing that leaching is controlled by external diffusion and chemical reactions.     

Atmospheric leaching characteristics of nickel and iron in limonitic laterite with sulfuric acid in the presence of sodium sulfite

Atmospheric leaching of nickel from limonitic laterite ores is regarded as a promising approach for nickel production, despite its low nickel recovery and slower leaching rate than high pressure acid leaching. Sulfur dioxide can enhance the sulfuric acid leaching of laterite, but its behavior for enhancing atmospheric sulfuric acid leaching was uncertain due to SO₂ losses and emission. In this study, sodium sulfite was used as a substitute for SO₂ gas in the leaching and the sulfuric acid leaching characteristics of Ni and Fe from a limonitic laterite in the presence of sodium sulfite were investigated. A linear correlation exists between the extraction of Ni and Fe, indicating the difficulty in selective leaching of Ni over Fe. Most nickel is isomorphically substituted within the goethite and it is difficult to dissolve in a high oxidation–reduction potential solution environment, resulting in a low nickel recovery. SO₂(aq) generated from the reaction of sodium sulfite in sulfuric acid solution, lowers the potential for the reducing reaction of FeOOH to give Fe²⁺, accelerating the iron extraction and nickel liberation from goethite.     

High pressure acid leaching of a refractory lateritic nickel ore

This paper describes the experimental findings of the extraction of nickel and cobalt by high pressure acid leaching (HPAL) of a refractory limonitic nickel laterite ore from the Gördes region of Manisa in Turkey. By optimizing the basic HPAL process parameters: leaching at 255 °C with 0.30 sulfuric acid to ore weight ratio with a particle size of 100% −850μ for 1 h of leaching, it was found that 87.3% of nickel and 88.8% of cobalt present in the ore could be extracted into the pregnant leach solution (PLS). However, these extraction results were found to be relatively low compared with other similar studies. In order to understand the possible reasons for this relatively lower extraction, further investigations have shown that together with a problem related to the kinetics of the dissolution reactions, a persistent acid resistant refractory mineral present in this sample also limited the leaching process. Attempts were made with different additives to solve this problem. The effects of chemical additives such as HCl, Na₂SO₄, FeSO₄, Cu⁺ and sulfur were tested and the effect of each addition on the degree of extraction of nickel and cobalt was determined.     

Calcination of low-grade laterite for concentration of Ni by magnetic separation

With the continuous depletion of high-grade nickel ores such as millerite and niccolite, nickeliferous laterites have become the major source for the production of nickel metal. However, only 42% of the world’s production of nickel comes from laterites, since the concentration of Ni is relatively low (ca. 2 wt.%). In addition, other metals, such as magnesium, iron and silicon can be found in laterite, which make the concentration of nickel even more difficult.In this study, a low-grade nickeliferous laterite ore was first calcinated and then processed by using a wet magnetic separator in order to recover nickel. Since, the ore contains both Ni and Fe, the calcination of laterite is effective in altering the crystalline structure of Fe species and therefore its magnetic properties, which in turn enable the selective concentration of nickel by magnetic separation that is an easy and environmentally-friendly technique. The experimental results have indicated the importance of carefully controlling: (1) the calcination temperature; (2) the pulp density and (3) applied magnetic field strength. The main finding of this work was that magnetic separation is effective in recovering 48% of nickel from laterite, increasing the Ni grade in the recovered product from 1.5% to 2.9%, when prior to the separation the ore was calcinated at 500 °C for 1 h.