H2SO4 Leaching Research Articles

Effect of H3PO4 coordination on vanadium extraction and iron separation in a H2SO4 leaching system.

The selectivity of the full wet leaching process for vanadium extraction using H2SO4 is low, resulting in a high impurity content in vanadium extracted from vanadium-bearing shale. This study focused on vanadium extraction and iron separation from vanadium-bearing shale, involving the coordination of H3PO4 in an H2SO4 leaching system. The effects of the ratio and quantity of H2SO4-H3PO4, leaching time, leaching temperature, and liquid-to-solid ratio on vanadium-bearing shale leaching were investigated. The dissolution processes of various minerals and the mechanism of iron coordination precipitation were analyzed. Results showed that vanadium leaching efficiency was 91.07% and iron leaching efficiency decreased from 84% to 23.86% under optimal leaching conditions-H2SO4-to-H3PO4 ratio of 2 : 1, H+ content of 8 mol kg-1, liquid-to-solid ratio of 0.8 L kg-1, and leaching time of 12 h at 95 °C. Leaching kinetics showed that the leaching process of vanadium shale was a mixed-control process in the H3PO4-H2SO4 leaching system; additionally, the leaching process was mainly controlled by a chemical reaction with an activation energy of 67 kJ mol-1. The preferential dissolution order of minerals in the vanadium-bearing shale was calcite, apatite, magnetite, muscovite, and pyrite. Under the H2SO4-H3PO4 leaching system, the iron content was reduced by inhibiting the dissolution of pyrite and coordination precipitation of Fe3+ with PO4 3-, thus separating iron and vanadium from the source. This provides guidance for vanadium extraction and impurity separation from vanadium-bearing shale using an all-wet method.

Efficient recovery of rare earth elements from ion-adsorption rare earth tailings: Based on the addition of pyrite calcination modification

Rare earth tailings constitute a significant amount of solid waste that remains after industrial (NH4)2SO4 leaching and typically contain rare earth elements (REEs), primarily Ce. This study aimed to fully recover REEs from rare earth tailings by calcination with pyrite under both air and N2 atmospheres, followed by dilute H2SO4 leaching or bioleaching. The results of XRD, SEM-EDS, TIMA, SEM-BSE, XPS, and ICP-MS analyses indicated that Ce and Mn were mainly present in the form of CeO2 and MnO2, respectively, whereas Fe existed in both Fe2O3 and Fe2(SO4)3 forms. Leaching experiments with dilute H2SO4 demonstrated that the co-calcination of rare earth tailings with pyrite significantly enhanced the recovery of REEs. Notably, calcination in air resulted in a significantly higher extraction rate of Ce (95.55 %) than that in N2 (83.82 %). This difference was attributed to the presence of O2, which promoted the oxidation of pyrite and facilitated the reduction of Ce(IV) to Ce(III). In contrast, incomplete oxidation of pyrite occurred in the N2 atmosphere, leading to a high residual Ce(IV) content in the tailings that could not be leached by the dilute acid. Subsequently, sufficient recovery of Ce (97.52 %) from the tailings calcined in the N2 atmosphere was achieved by bioleaching using Acidithiobacillus ferrooxidans and pyrite. Thus, this study provides theoretical support for the efficient recovery of REEs from mining tailings or secondary sources containing CeO2.

Preparation of V2O5 using Pb5(VO4)3OH recovered from Bayer vanadium sludge through H2SO4 pre-treatment, H2O2 leaching, V5O12∙6H2O and V10O24∙12H2O precipitation and oxidative roasting at 400 °C

Bayer vanadium sludge (BVS), discharged during the Bayer alumina production, is a promising secondary resource for the supply of vanadium. Previous vanadium extraction process involving Pb5(VO4)3OH precipitation by PbSO4, H2SO4 and (NH4)2SO3 leaching, (NH4)2S2O8 oxidation followed by ammonium vanadate precipitation, suffered from large amounts of reagent consumption and discharge of ammonia-containing waste. In this study, a cleaner and cost-efficient process based on H2SO4 pre-treatment, H2O2 leaching, hydrolytic precipitation, and preparation of V2O5 through oxidative roasting was proposed to reduce reagent consumption and waste. Results show that the maximum vanadium leaching efficiency reached 99.7 % under the most suitable conditions of H2SO4 pre-treatment with H2SO4/Pb5(VO4)3OH molar ratio of 6.5, pulp density of 250 g/L at 90 °C for 4 h followed by H2O2 leaching with H2O2/V molar ratio of 3.0 at 15 °C for 10 min. Furthermore, 96.5 % of vanadium was precipitated in the forms of V5O12∙6H2O and V10O24∙12H2O by heating the leach liquors to 100 °C for 60 min, followed by the oxidative roasting at 400 °C for 2 h to prepare V2O5 product with a purity of 99.2 %, with only H2O as a by-product. This research provides new insights into vanadium extraction in a cleaner, cost-efficient, and sustainable industrial process which saves 147.4 $/t Pb5(VO4)3OH.

Recycling of Ni from leached spent catalyst residue by H2SO4 leaching and solvent extraction: leaching kinetics, purification and product preparation

Recycling of Ni from leached spent catalyst residue by H2SO4 leaching and solvent extraction: leaching kinetics, purification and product preparation

Selective extraction of lithium and REEs from waste rare earth polishing powder: A two-stage leaching technique

Rare earth polishing powder waste (REPPW) is a Li-containing rare earth solid waste generated from glass-ceramic polishing processes. The key to its value-added recycling is selective extraction of Li and rare earth elements (REEs), however, only REEs recovery were addressed in conventional recycling processes. Aiming at this research gap, this work proposes a two-stage leaching technique consisting of Na2CO3 roasting, low concentration H2SO4 leaching (Li), and HCl reduction leaching (REEs). In the first stage, the leaching ratio of Li reached 99.0 %, while those of Ce and La were only 0.02 % and 2.5 %, respectively. It is found that inactive LiAlSi2O6 phase in the REPPW was transferred to leachable Li2SiO3 during Na2CO3 roasting with a leaching Ea of 85.32 kJ·mol−1, while La formed insoluble NaLa(SO4)2, whose precipitation route can be regulated by the leaching temperature. In the second stage, HCl-ascorbic acid (Vc) leaching system was adopted for REEs extraction from the first stage leaching residue, leading to an overall REEs recovery of 90.8 %. This work proposed a promising selective recycling route for Li and REEs in REPPW and would help for developing value-added REPPW recycling processes.

An efficient approach for preparation of battery-grade Li2CO3 from intermediate product Li2SiO3

The continuous development of lithium-ion batteries (LIBs) manufacturing industry calls for the mass production of Li2CO3, which serves as a pivotal raw material for LIBs. In our previous study, Li was successfully precipitated as intermediate product Li2SiO3 from an alkaline solution containing Si by the hydrothermal precipitation method, realizing the preliminary separation of Li and impurities. In this research, H2SO4 digestion-water leaching-Na2CO3 precipitation process was proposed to manufacture Li2CO3 from the intermediate product Li2SiO3. Firstly, the reaction feasibility of H2SO4 digestion-water leaching and direct H2SO4 leaching for Li2SiO3 were comparatively evaluated through thermodynamic calculation and experimental research. Compared to direct H2SO4 leaching, H2SO4 digestion-water leaching process is preferred in view of its excellent filtration properties. Subsequently, systematic digestion and leaching experiments showed that up to 97.1 % of Li was extracted under the optimized conditions, and the concentration of Li in the leach liquor is as high as 29.33 g/L. After pH adjustment and Al removal, Na2CO3 precipitation were adopted for Li2CO3 production. The equilibrium relationship in Li2SO4-Na2CO3-H2O system was analyzed to predict the precipitation behavior of Li and select appropriate operation window for Li2CO3 preparation. Finally, with an additional carbonation-decomposition process, a Li2CO3 product with a purity of 99.51 % is produced. This process realizes efficient synthesis of battery-grade Li2CO3 from the intermediate product Li2SiO3.

Harnessing Regenerated Graphite from Spent Lithium‐Ion Batteries to Enhance the Performance of Sulfur Cathode in Lithium‐Sulfur Batteries

AbstractIn this work, a facile hydrometallurgical approach combined with heat treatment has been adopted to recycle spent graphite from retired lithium‐ion battery cells. Graphite regenerated with 18 M H2SO4 leaching leads to a high specific surface area, significant porosity and small crystallite size. The regenerated graphite has been infiltrated with sulfur under solvothermal conditions at 160 °C. The sample containing 68 wt % of sulfur recovers a capacity of 224 mAh/g after 100 cycles. Furthermore, the electrode with a higher sulfur content of 84 wt %, recuperates the specific discharging capacity of 207 mAh/g after 100 cycles. After the initial fading during the first 5–10 cycles, the cathodes based on regenerated graphite demonstrate stable cycling behavior. This research not only offers an accessible and scalable method for regenerating graphite from spent LIBs but also demonstrates a new application of the regenerated graphite in LSBs, showcasing outstanding electrochemical performance.

Waste reduction and high value utilization of jarosite-alunite residue (JAR) produced from the recovery of spent lithium-ion battery

The residue known as jarosite-alunite (JAR) is produced when the leach solution of spent lithium-ion battery is neutralized. This residue typically consists of Fe, Al, Na2SO4, Ni, Co, and Mn. It is classified as both a hazardous solid waste and a secondary resource. A unique hydrometallurgical technique was implemented to recover Na2SO4 and use Al extracted from JAR in high value applications. This extraction process involves phase transformation and NaOH leaching, with the pH adjusted in the range from 10.9 to 14. Initially, the JAR compound underwent dissociation to isolate SO42− as Na2SO4 by means of NaOH at a moderate pH, while the other metals were preserved as a hydroxide residue. Afterwards, aluminum in the hydroxide residue was selectively leached with NaOH leaving Ni, Co, and Mn in the remaining residue. The results indicated that over 93% of Na2SO4 and 86% of Al in JAR were effectively recovered as Na2SO4 and high value-added γ-AlOOH, respectively. Additionally, the enriched Ni, Co, and Mn in the alkaline leach residue were selectively recovered by H2SO4 leaching. The suggested procedure led to a significant decrease in waste by more than 67%, offering a fresh approach to effectively reduce waste and recover metals from JAR.

Recycling and detoxification process of waste tire ash via acid leaching and neutralization

The possibility of a novel process of recycling and detoxification for waste tire ash, which contains hazardous Zn and Al, using acid leaching and neutralization was investigated. The waste tire ash mainly contains Zn, Al, Fe, Ca and Si. The leached contents of Zn, Al, and Fe with 1 mol/L of sulphuric acid (H2SO4) and hydrochloric acid (HCl) are almost same and the highest, while those of Ca with HCl is high and with H2SO4 is almost zero due to the formation of gypsum (CaSO4·2H2O). The optimum leaching condition is that waste tire ash was treated with 1 M H2SO4 for 1 h at 60 °C (solid-liquid ratio is 125 g/L). The elution of Zn and Al from the residue via H2SO4 leaching are 0.7 mg/L and 0.15 mg/L, respectively, which are lower than Japanese elution standard (Zn: 1 mg/L, Al: 0.2 mg/L), while those from raw ash are 3.3 mg/L and 1.4 mg/L. Zn-type layered double hydroxide (Zn-LDH) can be synthesized from leached liquor at 40 °C for 1 h by neutralization at pH 7. These results suggested that detoxified ash and environmental friendly material can be obtained from waste tire ash using H2SO4 acid leaching and neutralization.

Innovative recycling of high purity silver from silicon solar cells by acid leaching and ultrasonication

Precious and scarce silver (Ag) is used as a front electrical contact in silicon solar panels. With massive amounts of solar panel waste coming to end-of-life, it is imperative to recover all the Ag from these modules. In this paper, we propose a novel method to easily reclaim Ag from end-of-life silicon solar cells using low concentration sulfuric acid (H2SO4) leaching followed by ultrasonication. Our process simplifies the Ag recycling procedure by directly recovering the Ag contacts from solar cells, eliminating the need for secondary precipitation/electrodeposition. First, scanning electron microscopy (SEM) with energy dispersive X-ray spectroscopy (EDX) and X-ray photoelectron spectroscopy (XPS) was used to study the leaching and ultrasonication process on rejected solar wafers. Next, a solar panel from landfill was heated in a furnace to burn off the polymer encapsulant. The silicon wafers were then collected from the burning process and leached in 2 M H2SO4 at two different shaking speeds for 48 h at room temperature. Finally, the end-of-life silicon wafer pieces were collected, sonicated in water, and then the sonication water was centrifuged to show a proof-of-concept to recover the Ag contacts. Inductively coupled plasma optical emission spectroscopy (ICPOES) is used to monitor dissolved elements in the leachate as a function of time and shaking speed. XPS is used to evaluate the composition of the silicon cell wafer surface before and after H2SO4 leaching. SEM is used to image the recovered Ag contacts morphology and EDX confirms the recovered particles contain high amounts of Ag. A proposed schematic illustrates the authors’ hypothesis for the peeling mechanism.

Observations and simulations for separation mechanism of vanadium and chromium: Progressive study based on oxides, spinels and vanadium-chromium slag

In this study, V2O3-CaO-Cr2O3 oxides system, Fe2VO4-CaO-FeCr2O4 spinels system and V-Cr slag-CaO system were established progressively to investigate the separation effect of vanadium and chromium components in vanadium-chromium slag (V-Cr slag). Then, the separation mechanism was proposed by XRD, SEM and leaching experiments. The results show that the reaction ability of vanadium with calcium is far stronger than that of chromium. In oxides and spinels systems, as CaO increases, CaV2O6, Ca2V2O7, and Ca3V2O8 are generated sequentially. In V-Cr slag-CaO system, additive CaO, along with manganese, magnesium and iron components in V-Cr slag reacts with vanadium, resulting in the complex composition of vanadate. As to chromium component, Cr2O3 in oxides system will be converted into CaCrO4 as n(CaO)/n(V2O3) is higher than 2. Iron in the spinels system reacts with chromium and form solid solution (FexCr1-x)2O3, resulting in less chromium reacting with calcium to produce CaCrO4. In V-Cr slag-CaO system, SiO2 generated by the decomposition of silicate phase combines with part of calcium expected to react with chromium, making CaCrO4 generation more difficult, promoting the separation of vanadium and chromium further. Consequently, during H2SO4 leaching, chromium leaching ratios are 98.75%, 45.77% and 5.91%, respectively, in oxides, spinels and V-Cr slag system when n(CaO)/n(V2O3) is 5. And when n(CaO)/n(V2O3) is 2, more than 90% of vanadium can be leached, however, the leaching ratios of chromium are 0.22%, 0.17% and 0.094% in these three systems. Furthermore, the variation regularities of chromium leaching ratios with CaO amount are described by regression equations. These findings would provide the theoretical support for the calcification roasting process with V-Cr slag.

Sodium carbonate roasting and mild acid leaching of vanadium titanomagnetite concentrates: Vanadium extraction and residue sodium decrease

Vanadium titanomagnetite is a crucial resource for vanadium and titanium. The traditional process of treating vanadium titanomagnetite concentrates has shortcomings like low vanadium recovery rate and titanium recovery inability, resulting in resource wasting and solid waste discharging. Thus, a new process of Na2CO3 roasting and mild H2SO4 leaching has been proposed. In the concentrates, titanium occurs majorly in titanomagnetite (Fe3-xTixO4, 0≤x≤1) and ilmenite (FeTiO3), vanadium in titanomagnetite and sphene (CaTiSiO5), calcium in sphene, and silicon and aluminum in several minerals. Phases are tuned during Na2CO3 roasting, where V and impurities Al and Si transform into soluble sodium compounds, leachable phase sphene doesn't change, and Ti is enriched in insoluble phases. Then, leached by mild H2SO4, the soluble compounds are almost all leached, and the leaching rates of V, Ca, Al, and Si are 91%, 76%, 84%, and 82%, respectively. The Na2O content in the leaching residue decreases to 1.0 wt%, which enables the residue for Ti extraction. This process is expected to recover V and Ti comprehensively from the vanadium titanomagnetite concentrates. The roasting kinetics are also studied based on CO2 content in the off-gas for a deeper understanding of the mechanism.

An Environmentally Friendly Sulfuric Acid Decomposition Strategy for Mixed Rare Earth Concentrate

A novel environmentally friendly one-step decomposition strategy for mixed rare earth concentrate of Bayan Obo in sulfuric acid solution was proposed in this work. In this process, more than 84% of bastnasite and monazite were decomposed in the leaching step at a temperature lower than the boiling point of sulfuric acid solution. So, the dilapidation of sulfuric acid in this current proposed process will be reduced to a large extent. The stability region of rare earth ion in the RE(La, Ce, Nd)-F-P-SO4-H2O system at 170 °C has been proven through Eh-pH diagrams. The factors influencing decomposition of rare earth concentrate in this process were also investigated and the optimal leaching conditions were determined to be a leaching temperature of 170 °C with an ore/acid ratio of 1:5 (g/mL), a sulfuric acid concentrate of 75% and a leaching time of 80 min. The mineralogical changes occurring during the H2SO4 leaching process were investigated by X-ray diffraction and SEM-EDS. The analysis results showed that bastnasite and most of monazite had been decomposed, leaving only a small amount of monazite in the leaching residue.

Study on Extraction Valuable Metal Elements by Co-Roasting Coal Gangue with Coal Gasification Coarse Slag.

Coal gangue (CG) and coal gasification coarse slag (CGCS) possess both hazardous and resourceful attributes. The present study employed co-roasting followed by H2SO4 leaching to extract Al and Fe from CG and CGCS. The activation behavior and phase transformation mechanism during the co-roasting process were investigated through TG, XRD, FTIR, and XPS characterization analysis as well as Gibbs free energy calculation. The results demonstrate that the leaching rate of total iron (TFe) reached 79.93%, and Al3+ achieved 43.78% under the optimized experimental conditions (co-roasting process: CG/CGCS mass ratio of 8/2, 600 °C, 1 h; H2SO4 leaching process: 30 wt% H2SO4, 90 °C, 5 h, liquid to solid ratio of 5:1 mL/g). Co-roasting induced the conversion of inert kaolinite to active metakaolinite, subsequently leading to the formation of sillimanite (Al2SiO5) and hercynite (FeAl2O4). The iron phases underwent a selective transformation in the following sequence: hematite (Fe2O3) → magnetite (Fe3O4) → wustite (FeO) → ferrosilite (FeSiO3), hercynite (FeAl2O4), and fayalite (Fe2SiO4). Furthermore, we found that acid solution and leached residue both have broad application prospects. This study highlights the significant potential of co-roasting CG and CGCS for high-value utilization.

Facile manufacture of high-purity CuSO4 from waste Cu-containing paint residue using combined processes of H2SO4 leaching and extraction stripping

Abstract Waste copper-containing paint residue (WCPR) represents a typical hazardous waste containing both toxic organic substances and toxic heavy metals, but there are few reports on the recycling of heavy metals. The recovery of Cu from WCPR by H2SO4 leaching–extraction–stripping has the advantages of eco-friendliness, simplicity of operation, and high value-added product. The results show that under the optimal conditions, the leaching rate of Cu in WCPR is 94.31% (18.02 g/L), while the extraction and stripping rates of Cu in the leaching solution are 99.46 and 95.32%, respectively. Due to the high concentration of Cu2+ with fewer impurities in the stripping solution, the stripping solution is heated, evaporated, cooled, and crystallized to successfully produce high-purity dark blue CuSO4 crystal, accomplishing the high-value recycling of Cu in WCPR. In addition, the leach residue of WCPR contains acrylic resin and SiO2, which can be used in cement kilns for incineration, thus realizing the overall recycling and utilization of WCPR.

Innovative Recovery of Gallium and Zinc from Corundum Flue Dust by Ultrasound-Assisted H2SO4 Leaching

ABSTRACT Rare metal gallium is classified as a strategic material because of continued high demand, uncertain supply, and low reserves. Corundum flue dust (CFD) is generated during bauxite smelting to produce brown corundum. Gallium (Ga), Zinc (Zn), and potassium in the bauxite could be enriched with CFD through this process, so CFD is regarded as one of the potential sources of Ga and Zn. A novel that utilizes ultrasound to enhance the H2SO4 leaching efficiency of Ga and Zn from CFD was developed from CFD. The effects of ultrasound power, leaching time, sulfuric acid concentration, and leaching temperature on the leaching behavior of Ga and Zn were studied, and the corresponding kinetics model was established. It was found that the maximum Ga and Zn leaching efficiencies of 81.22% and 96.72% were obtained at optimum leaching conditions: the leaching temperature of 90°C, the leaching time of 60 min, 3 M H2SO4, the solid-to-liquid ratio of 1:5 (g/mL), and the ultrasound power of 750 W. In all cases, the leaching effect of Zn is not significantly improved by ultrasonic treatment, but the leaching efficiency of Ga has a positive effect (more than an 18% increase). Kinetic analysis shows that the leaching process of Ga and Zn were controlled by chemical reaction and diffusion, respectively. The apparent activation energy values are 20.39 kJ/mol for Ga and 13.44 kJ/mol for Zn. The characterization of the leaching residues reveals that the ultrasound effected to enable the whole block to be fractured or dispersed into fragments, accelerating the penetration of H2SO4 into the pores and cracks at the solid, which further prompts the dissolution of the encapsulated gallium and achieves the conspicuous improvement to Ga leaching efficiency.

Optimisation of extraction of valuable metals from waste LED via response surface method

AbstractThe valuable metals were separated and extracted from waste light-emitting diodes (LEDs) by pyrolysis and leaching process. About 98.16% Ga and 99.54% Y were recycled from the LEDs after the pyrolysis by sieving process in advance. The effects of various leaching conditions on the extraction of Fe, Cu and Ag were investigated by single factor experiments and response surface method. About 90.15% Fe was selectively extracted from the waste LEDs by H2SO4 leaching without oxidant, and 99.55% Cu and 99.36% Ag were recovered from the iron leaching residue by acid leaching with the addition of HNO3. The XRD and SEM−EDS analyses confirmed that the iron and copper were effectively extracted from the waste LEDs by the developed process.

Removal of Pyrrhotite from High-Sulphur Tailings Utilising Non-Oxidative H2SO4 Leaching

Tailings are a residual material stream produced in the mineral processing of ores. They may contain a major sulphide content that increases the risk of acid rock drainage (ARD) but may also host valuable metals. Tank bioleaching is a technically viable method to treat sulphide tailings. However, a significant pyrrhotite content may cause increased acid and oxidant consumption and result in longer retention times in a bioleaching process. In this work, non-oxidative H2SO4 leaching of pyrrhotite is studied for high-sulphur tailings, both as a pre-treatment method and to consider the recovery possibilities of Fe and S. Continuous mode validation tests, conducted at 90 °C, pH 1.0 and 106 min retention time, resulted in a complete pyrrhotite dissolution with 427 kg/t acid consumption (as 95% H2SO4). Unwanted dissolution of Ni and Zn was taking place with a leaching yield of 21.5% and 13.5%, respectively, while Co and Cu dissolution was negligible. The continuous mode tests signalled that by shortening the retention time, Ni dissolution could be dramatically decreased. The non-oxidative pyrrhotite leaching produced a H2S-rich gas stream, which could be utilised in later metals’ recovery processes after bioleaching to precipitate (CoNi)S, ZnS and CuS products. The non-oxidative pyrrhotite leaching also produced a FeSO4 solution, with approximately 20 g/L of Fe.

Recovery of arsenic oxide, harmless gypsum residue and clean water by lime neutralization and precipitation

The nonferrous metal smelting industry-an influential industry in China- produces a large quantity of arsenic-containing acidic wastewater (ACAW) that seriously hinders the development of the industry. In this study, low-cost and effective methods for treating acidic wastewater were systematically examined. Lime neutralization and compound salt precipitation were used to remove arsenic (As) from acidic wastewater. The residues produced during neutralization and compound salt precipitation were treated harmlessly by H2SO4 leaching, and As2O3 was recovered from the sulfuric acid leach solution (ALS). The results showed that the appropriate pH for the neutralization process was 3. The arsenic concentration in the solution after compound salt precipitation was <0.5 mg/L at pH 8.5 and suitable molar ratios (Cu/As: 0.45 to 2.3, Fe/As: 1.20 to 6.0, and Zn/As: 1.20 to 6.0). After acid leaching of residues obtained during neutralization and compound salt precipitation, the content of CaSO4 in the leach residue reached the specified value (CaSO4 ≥ 75 wt%) of industrial byproduct gypsum (GB/T 21371–2019 of China), and the concentration of arsenic in the toxicity leaching experiment was <5 mg/L. The ALS was reduced by SO2 to recover arsenic trioxide, and the recovery of arsenic reached 96.5%. Compared to the NaOH neutralization reagent, the lime treatment process is cost-effective and produces non-hazardous solid waste.

Production of Al(HI)-K(I)-Ti(IV)-sulphate-containing leach liquor from metakaolinite-containing ash derived from South African coal fines

South African discard coal fines and K2CO3 blends were heated in a laboratory-scale rotary kiln to produce ashes for H2SO4 leaching tests. The optimized H2SO4 leaching conditions of 6.12 mol.dm3 (M) H2SO4, solid to liquid ratios 1:5 and 1:10, and 8o°C for 8 hours were used. K2CO3 was added to increase the dissolution efficiency of K. The objective was to determine if the Al present in metakaolinite (Al2O3.2SiO2), the Al, K, and Ti in the alumino-silicate glasses, and the Ti in rutile (TiO2) in the ashes could be selectively dissolved in H2SO4. XRF results show that the ashes formed at 700°C dissolved more efficiently (87% Al, 89% K and 23% Ti) compared to the ashes formed at 1050°C. This can be attributed to the presence of Al2O3.2SiO2, K2CO3 melt, K2CO3 remnants, KAl(SO4)2, and K-aluminosilicate glass in these ashes. XRD results indicate that the ashes prepared at 1050°C contained anorthite (CaAl^Og), microcline (KAlSi3O8), pseudomullite (Al2.SiO2), and silicon spinel (2Al2O3.3SiO2), which are either insoluble or only sparingly soluble in H2SO4. These minerals resulted in the lower dissolution efficiencies of Al and K. Based on the high dissolution efficiencies of Al and K for the ashes produced at 700°C, coal fines blended with K2CO3 could possibly be utilized as feedstocks for the production of aluminium(III), potassium(I), and titanium(IV) and a sulphate-containing leach liquor. Furthermore, the environmental issues and costs associated with the handling and disposal of large volumes of coal fines will also be resolved.