Browsing by Author "Mokone, Thebe"

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    Direct oxidative Ammonia-Ammonium salt leaching of sphalerite concentrate from the Gamsberg Mine and study of the catalytic effect of Cu(II) on Zinc extraction
    (2025) Shabalala, Sanele; Petersen, Joachim; Mokone, Thebe
    The feasibility of using ammonia-ammonium salts to selectively leach Zn over Fe from Gamsberg sphalerite flotation concentrate under moderate temperatures and atmospheric pressure was investigated in stirred tank reactors. This approach is suggested as an alternative primary method for extracting Zn from sphalerite concentrates that are not environmentally friendly to process with conventional sulfuric acid leaching techniques. The analysed concentrate consists of 43.7 wt.% Zn, 0.17 wt.% Cu, 8.3 wt.% Fe, 29.1 wt.% S, 3 wt.% Pb, 2.4 wt.% Mn, and 5.4 wt.% Si, with mineralogical components including sphalerite (81.2 wt.%), galena (3.2 wt.%), quartz (9.3 wt.%), mica (2 wt.%), clinochlore (1.7 wt.%), and plagioclase (2.6 wt.%). This study begins with preliminary tests to evaluate the effectiveness of ammonium salts (chloride, sulphate, and carbonate) and to examine the catalytic effect of Cu(II) ions on the direct selective leaching of Zn over Fe from the concentrate, using oxygen and compressed air as the oxidative mediums. Ammonium chloride was the most effective salt for selectively leaching Zn over Fe, and the addition of Cu(II) ions improved Zn extraction in an oxygen medium. Ammonium chloride was then used to optimise parameters such as temperature, total ammonia concentration, pulp density and Cu(II) concentration for Zn extraction. The optimal conditions found were 6 mol/L [NH3]T (NH3/NH4+ ratio of 1), ~0.38 g/L Cu(II) concentration, 2% w/v pulp density, a leaching temperature of 55°C, agitation of 500 rmp in oxygen for 72 hours. Under these conditions, a Zn extraction efficiency of 99.5% was achieved. The rate of Zn extraction increased with rising temperatures (35–55°C), but at higher temperatures (75°C), NH3 evaporation impacted the initial extraction rates. The Zn extraction rates also increased with an increase in total ammonia concentration and Cu(II) concentration but decreased with an increase in pulp density. Initial slope analysis was used to characterize the leaching kinetics for each particle, providing good linear fits (R2 > 0.99) across all temperatures, ammonia concentrations, Cu(II) concentrations, and oxygen partial pressures. The activation energy was calculated to be 41.6 kJ/mol using the Arrhenius equation, suggesting that the leaching process was controlled by a surface chemical reaction and was highly sensitive to temperature. The shrinking core models were plotted under optimal conditions, confirming that the leaching process was governed by surface chemical reaction control. The reaction order for total ammonia concentration, Cu(II) concentration, and oxygen partial pressure were determined to be 0.22, 0.079, and 0.46, respectively. Due to its high volatility, NH3 evaporates rapidly. The concentration of total evaporated ammonia was determined through spectrophotometric analysis by measuring the total ammonia accumulated in scrubber bottles from the reactor after a set leaching period. The concentration of evaporated NH3 was found to increase with longer leaching times, higher temperatures, and greater initial total ammonia concentrations, with the highest evaporation observed at 75°C after 3 days of leaching. For optimal reagent conservation, a leaching temperature of 55°C is recommended for ammonia leaching of sphalerite from the Gamsberg mine.
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    Vanadium and titanium extraction from titaniferous slag using a roast-leach process
    (2025) Nkosi, Sanele; Mokone, Thebe; Petersen, Jochen; Goso, Xolisa
    Vanadium (V) is typically extracted from the titaniferous magnetite (titanomagnetite) ore using either the V primary production and V and steel co-production process. The V primary production process is the roast-leach process which involves roasting, leaching, precipitation, and calcination of the titanomagnetite. The V and steel co-production process entails the smelting of titanomagnetite in the presence or absence of the fluxes. In the presence of flux i.e. the fluxed smelting approach, most of the V reports to the pig iron, whilst titania (TiO2) reports the slag. This slag is discarded since the TiO2 grade is considered too low for upgrading. Also, this titaniferous slag has a complex phase chemistry such that the spinel and pseudobrookite solid solution (ss) phase incorporate most of the V and Ti species. The objective of this project was to investigate a roast-leach process that would maximize the extraction of V from the titaniferous slag and remove high amounts of impurities from the water leach residue to maximize the TiO2 grade in the product residue. The slag produced by the now-defunct EVRAZ Highveld Steel and Vanadium Cooperation (EHSV) of South Africa (SA) was used as a case study. The EHSV titaniferous slag contains about 0.9% V2O5 and 35.6% TiO2. The best roasting conditions were investigated through the variation of Na2CO3: NaOH ratios, stoichiometric Na additions, roasting temperatures, roasting times, particle size distribution (PSD), and Na reagents i.e. Na2CO3, NaOH, and Na2SO4 salts. The roasting stage is aimed at the conversion of V2O5 in the titaniferous slag to a water soluble sodium metavanadate (NaVO3). The produced roast products were leached using water at standard conditions of 70°C, 120 minutes leaching time, s:l ratio of 1:4, and agitation speed of 350 rpm. The produced water leach residue was further subjected to acid leaching using HCl as a lixiviant with 20% acid concentration, at 110°C, 24 hours, s:l ratio of 1:4, and agitation speed of 350 rpm. The acid leaching stage was conducted in order to remove impurities such as Al, Ca, Mg, and Fe. The resulting acid leach residue was further upgraded through caustic leaching for Si impurity removal. The standard caustic leaching conditions that were used include 2.15 M NaOH solution as a lixiviant, leaching temperature of 100°C, leaching time of 3 hours, s:l ratio of 1:4, and agitation speed of 350 rpm. The best roasting conditions were used to measure the best leaching conditions by variation of the leaching time from 1 to 24 hours and acid concentration from 15% to 25%. The residue from the best leaching conditions were subjected to standard caustic leaching conditions. The high TiO2 grade that resulted from the removal of impurities was washed to remove the Na present in the residue. The best roasting conditions were 200% stoichiometric Na addition of Na2CO3: NaOH ratio of 100:0 and 0:100, 1000°C, 120 minutes, PSD of -850+105 μm. The best V extractions of 75.7% and 73.7% were attained when 200% stoichiometric Na2CO3 and NaOH were used respectively. Na2CO3 salt was the reagent that was used for downstream upgrading of the water-leach residue since it is cheap and used in industrial operations. The TiO2 grade of 68.7% was attained in the caustic-leach residue when the best roasting conditions were used for roasting the titaniferous slag. The phase composition results showed that pseudobrookite ss and the spinel do decompose when excess Na was added. After this decomposition, the SEM showed that the Na3MgAlSi2O8 phase forms after roasting. This phase also decomposed during water leaching to form the Na2MgTi2O6 phase which contains high Ti and Si species and several impurities. Acid leaching results showed the possibility of minimising Mg, Ca, and Fe impurities. Al impurities in the acid-leach residue remained high in concentration. The phase composition results showed that Al species were present in all the phases that formed after acid-leaching. The caustic leaching process was a success since most of the Si impurities were removed. The phase composition results showed that the minor Si species present in the caustic-leach residue were in the rutile phase. The produced caustic-leach residue still contained several impurities. The best leaching conditions were 24 hours leaching time and 25% HCl lixiviant concentration. Using these best conditions resulted in TiO2 grades of 71% in the caustic leach residue. Washing of this caustic leach residue further increased the TiO2 grade to 79.6%. The phase composition of the washed sample showed that the concentration of Na and Si impurities in the residue decreased. The rutile phase also contained a variety of Mg, Al, Cl, and Ca impurities. The impurity extraction degrees showed an increase with increasing leaching time. The extraction degrees followed the order of Mg > Fe > Al > Ca when leaching times were varied. The R2 values of Mg and Fe showed that the rate controlling mechanism is the surface chemical reaction whilst Al and Ca are controlled by diffusion across product layer and interface transfer. The evaluation of V and impurity extraction from the titaniferous slag was possible at the attained best conditions. The impurities were not fully removed to their maximum, hence the high TiO2 grade product was still contaminated with impurities. The produced TiO2 grade did not meet the TiO2 feedstock specifications as the MgO, CaO, Al2O3, and FeO impurities were still high in the final product, therefore further optimization work is required.