Aleks Nikoloski

Conventional rotary kilns used in spodumene decrepitation by calcination have difficulties in the processing of fine spodumene concentrates. Fine particles are more susceptible to melting in the kiln, rendering the lithium unrecoverable. The loss of fines as dust is another potential problem. Processing ores in which the spodumene is more disseminated, and the use of flotation to concentrate spodumene, results in finer grained concentrates. It is therefore necessary to develop alternative processes that can handle fine grained spodumene concentrates.

One alternative is flash calcination, where the material freefalls through a vertical shaft kiln. The grains are separated, and the α→β spodumene transition occurs rapidly during the descent. A spodumene concentrate containing 6.0%, with a size range of 90% passing 200 μm, was calcined in a Calix reactor, which is a new type of flash calcination kiln. Rapid conversion of α-spodumene to β-/γ- spodumene was achieved using this new furnace, though multiple passes were needed to achieve good conversion percentages. Four passes at 1050°C resulted in 54% conversion, four passes at 1100°C resulted in 88% conversion, and two passes at 1120°C resulted in 84% conversion. When two different size fractions, -106 μm and +106 μm were treated under the same conditions (one pass, 1100°C) there was minimal difference in the extent of conversion.

Acid baking followed by water leaching of the calcined samples was run under a standard set of conditions: 180% stoichiometric acid requirement, 250°C for 1 h followed by 2 h of leaching in water at 50°C. Lithium extractions correlated closely with the extent of spodumene conversion of the calcined samples as measured by chemical and XRD methods.

Brannerite, UTi2O6, is the most important uranium mineral after uraninite and coffinite, and the most common refractory uranium mineral. As the more-leachable ores become exhausted, it is necessary to process complex and refractory ores to meet the growing demand for uranium as an energy source. The present study was carried out to provide information that will enable the development of a more effective processing strategy for the extraction of uranium from ores containing brannerite. A detailed study was carried out to understand the leaching behaviour of brannerite in sulphate media (10–200 g/L H2SO4) under moderate temperature conditions (25–96°C), and in alternative acid and alkaline systems. The feed and the leached residues were characterised by X-ray diffraction and scanning electron microscopy with energy-dispersive spectrometry techniques. The brannerite dissolved up to 95% after 5 h leaching in ferric sulphate media, up to 89% in ferric chloride media under similar conditions, and up to 82% in 24 h in sodium carbonate media. The alkaline leaching was repeated with a high-carbonate brannerite-bearing ore, which showed comparable extractions. Mineralogical characterisation showed that altered and amorphous regions are a regular feature of brannerite, and that pitting is typically observed on the surface of the leached grains. Brannerite was shown to generally dissolve congruently, with altered and amorphous regions in the brannerite grains dissolving faster than the crystalline regions, which implies that the extent of brannerite alteration is a key parameter in the process selection, along with the grade, liberation size and gangue mineralogy.

Brannerite, UTi2O6 is the most important uranium mineral after uraninite, UO2 and coffinite, U(SiO4)1-x(OH)4x. It is also the most common refractory uranium mineral. Ores containing brannerite typically require intense conditions (>50 g/L H2SO4, >75°C) compared to other uranium ores for effective uranium extraction to occur. To develop an effective process for the extraction of uranium from brannerite containing ores and improve the extraction from the ores currently being processed, it is necessary to understand the chemistry of the brannerite leaching process. As this study has shown, brannerite is typically an altered and amorphous mineral, with an extent of alteration depending on the age of the sample and the geological history of the deposit. A sample of brannerite from Cordoba, Spain, was leached over a range of conditions in acidic ferric sulphate media. The sample was filled with cracks and altered zones containing anatase (TiO2). Process parameters studied included temperature (25-96°C), acidity (10-200 g/L H2SO4) and the effect of adding selected gangue minerals (apatite, fluorite and ilmenite). The feed and the leached residues were characterised in detail by XRD and SEM-EDX techniques. The results of this study showed that brannerite dissolution has a stronger dependence on temperature and lesser dependence on free acid concentration. Comparisons between the residues and the feed showed that the altered and amorphous areas of the brannerite sample are more readily leached than crystalline areas. The crystalline areas of the brannerite dissolved congruently, with titanium subsequently precipitating as anatase physically separated from the original brannerite grains.

Thiosulphate has received much attention as an alternative non-cyanide lixiviant for gold recovery over the last three decades. In particular, a number of studies have shown that an ammoniacal copper(II)/thiosulphate system offers fast leaching kinetics, but there are difficulties in controlling the complex solution chemistry and there are concerns over the use of ammonia. Recently, thiosulphate leaching of gold in the absence of ammonia has shown to be one of the most promising alternatives to cyanide, as evident from the thiosulphate gold processing plant recently commissioned at Barrick Goldstrike for treating pressure-oxidized double refractory ore. However, the published information on non-ammoniacal thiosulphate systems is limited. In this work, the dissolution of gold in non-ammoniacal thiosulphate solutions has been studied using a rotating electrochemical quartz crystal microbalance (REQCM), rotating gold disk, gold powder, and selected sulphidic gold ores. The electrochemical studies found that gold oxidation is enhanced by increases in temperature, thiosulphate concentration, and the addition of low levels of copper. Oxygen reduction was found to occur much more readily on sulphide mineral surfaces than on the gold surface, offering an opportunity for galvanic interaction. The subsequent leaching tests using REQCM showed that the gold leach rate in the oxygen-thiosulphate system without any additives is in the order of 10-7 mol m-2 s-1 , two orders of magnitude lower than a typical cyanidation rate. However, the use of elevated temperature, high oxygen concentration, and copper addition, in conjunction with the galvanic effect of sulphide minerals, dramatically improved the gold leach rate to the same order of magnitude as a typical cyanidation rate. This was supported by the results obtained from prolonged leaching tests using gold powder and sulphidic gold ores. This study hence shows that the oxygen-thiosulphate system could be a promising alternative to cyanidation for treating some sulphidic gold ores.

Metal oxide matrix composite anode containing a coating of lead and cobalt was produced and tested to develop an improved yet low-cost alternative anode for base metal electrowinning. The novel anode consists of a top coating of PbO2-Co3O4 electrodeposited onto Ti substrate with a protective interlayer of SnO2-Sb2O3. The anode was tested under typical copper electrowinning conditions. The results showed that the novel anode possesses considerable advantages compared to the conventional PbCaSn type anodes, including catalytic activity for oxygen evolution, reflected in up to 300 mV lower anode potential resulting in corresponding energy savings, and significantly lower corrosion rates resulting in lower operating costs and higher quality cathode. Continuous anodisation over a one week period showed a corrosion rate of approximately 0.1 g m-2 h -1. Cross section studies showed that the top coating of PbO2-Co3O4 and interlayer of SnO2-Sb2O3 effectively stopped the penetration of acid to the surface of the Ti substrate during anodisation, preventing passivation. Moreover, the lower corrosion rate enabled the production of virtually lead-free cathodes.

Metal matrix and metal oxide matrix composite coatings containing lead and cobalt were electrodeposited onto PbCaSn and Ti substrates in an effort to develop an improved yet low-cost anode for base metal electrowinning. The anodes were tested under typical copper electrowinning conditions to study their polarisation behaviour. Lead balance method was used to calculate the corrosion rates of composite coated anodes and compare this with the corrosion rate of a conventional uncoated PbCaSn anode. The results show that incorporation of cobalt into the surface of lead has the effect of reducing the anode potential and corrosion rate during 7-day tests. Some of the new anodes showed much better performance than PbCaSn in 168 hour tests.

A novel method for preparation of iron oxyhydroxide materials, involving aqueous precipitation followed by microwave assisted aging is investigated. The produced materials are characterized by XRD, SEM EDX and TEM spectroscopy and BET analysis. The materials show physical characteristics dependent on preparation procedure. The arsenic adsorptive properties of the materials are studied by batch adsorption techniques. It is found that the rate of arsenic upload depends strongly on the degree of crystallinity of the materials. The adsorption capacity is approximately 55 mg/g. The physical characterization of the arsenic loaded adsorbents shows that the adsorption process modifies the morphology of the materials. Over 4% of arsenic atoms are incorporated into the particle matrix.

Microfluidics is concerned with the characteristics and manipulation of fluid flows along mili- and micro-sized channels at very low Reynolds numbers intended for specific applications. It affords improvements based on several factors such as enhanced surface area-to-volume ratio and faster diffusion which facilitate efficient extractions and phase separations. In this study, microfluidic solvent extraction (μSX) was investigated as an alternative to conventional bulk solvent extraction of copper from a mineral leach solution. While bulk extraction typically requires a two-step mixer-settler stage, μSX precludes the need of a settler. A copper solution was produced by using sulfuric acid as leachant of a composite malachite–chalcopyrite ore. A microfluidic Y-Y channel placed on the format of a microscopy polymethylmethacrylate (PMMA) slide was used for the μSX process. LIX84 (10% v/v) dissolved in Shellsol was used as the extractant. In a typical μSX application, two streams (one is the copper leach solution and the other is the extractant) are in direct parallel contact along the microsized channel. There was no observable blockage along the microfluidic channel and copper extraction was accomplished under continuous phase flow.

Ion exchange resin has been used to recover value metal in the uranium and gold industry through resin-in-pulp/leach and similar carbon-in-leach/pulp processes for decades. More recently, resin-in-pulp processes have gained attention as a potential method to improve the efficiency of nickel operations. While every nickel laterite operation is unique, many involve an acid leach, neutralization and oxidative precipitation of impurities followed by counter current decantation to separate valuable liquor from the unwanted metal residue and precipitate. Counter current decantation (CCD) of this material is challenging at best, with large CCD tanks having a large plant footprint and requiring high capital investment. Depending on the settling characteristics of the precipitate, 5% or more of the leached nickel and cobalt can be lost to the slurry underflow through solution entrainment, co-precipitation, and sorption processes on the high surface area solids. For a site producing 40,000 tonnes per annum nickel and 2,500 tonnes per annum cobalt this represents yearly losses of approximately $40 million USD, given current LME spot prices (as of March 2012). The tremendous waste of value that these high losses of nickel and cobalt represent are the primary driving force behind the development of resin-in-pulp (RIP) scavenging from laterite tailings. RIP scavenging involves contacting ion exchange resin with nickel laterite tailings at conditions where the valuable metals load onto the resin. As the resin beads are larger than the slurry particles, they can be separated from the slurry using vibrating sieving. Following this, the resin is washed to remove residual slurry and solution, and then eluted to recover metal value. While exact values vary, typical caron process tails contain roughly 300 mg/L nickel and 50 mg/L cobalt in slurry. High pressure acid leach tailings may contain 200 mg/L nickel and 35 mg/L cobalt in slurry. With efficient resin-in-pulp contact, upwards of 90% of this otherwise lost metal value can be recovered. Although the chelating ion exchange resins proposed for use in nickel laterite RIP are selective for nickel and cobalt over other unwanted metals, laterite tailings solutions contain a relatively small amount of these metals of interest. Depending on the composition of the original ore and the method of leaching, the neutralized slurry can contain large amounts of solution phase magnesium and manganese (in the case of acid leaching) and vast amounts of ferric iron, silica, aluminium, and chromium in the solid phase. The presence of other cations that compete with nickel and cobalt for resin loading sites complicates resin-slurry equilibria. In general, there is a trade off between recovery of nickel and cobalt and purity of loaded resin. To recover a high amount of the nickel and cobalt value, one must accept the presence of impurity metals on the resin. When resin is eluted, these impurity metals can follow value metals into the eluate. To date, the majority of resin elution work has focused on metal recovery via acid contact (usually H2SO4). When one has produced a resin loaded with a high fraction of value metals, quantitative elution in this fashion is attractive. Using strong acid, metal is recovered in a small volume of eluent with rapid kinetics. However, as more impurities are loaded onto a resin, strong acid elution becomes less attractive as quantitative elution of a low purity resin produces a low purity eluate. In such a case, a method of selectively recovering value metal from resin is desirable. One selective elution method involves two stages – dilute acid to remove weakly bound impurity metals, followed by strong acid to recover the remaining metals[3]. For the iminodiacetic acid resins most commonly investigated for nickel RIP, this method achieves selectivity of nickel and cobalt over magnesium, calcium, and manganese, but does not separate nickel and cobalt from ferric iron or chromium. Another option is the use of ammoniacal elution[8]. Nickel and cobalt readily form stable amine complexes and have high solubility in strong ammonia solutions, unlike the majority of other metals present in nickel laterite processing. By taking advantage of this chemistry, nickel and cobalt can be effectively eluted separate from impurity metals. The value of selective elution depends on the purity of the loaded resin. With less impurities, the additional expense of a second stage of elution or the higher cost of ammonia reagents relative to sulphuric acid make selective elution less attractive. In order to better determine the performance of different elution methods, a resin produced through scavenging RIP of plant tails from Queensland Nickel was treated using three elution strategies – single stage strong acid, two stage weak acidstrong acid, and single stage ammoniacal elution.