This paper presents the study on recovery of rubidium (Rb), caesium (Cs), and potassium (K), elements of economic importance, from the leach residue obtained after sulphation roasting-water leaching (SRWL) of a lithium-bearing pegmatite ore via chlorination roasting-water leaching. The leach residue containing Rb 0.41 wt%, Cs 0.11 wt%, and K 4.14 wt% serves as the feed material. Rubidium and caesium partitioned into newly crystallized potassium aluminosilicate phases during the prior SRWL stage. The extraction was tested using different chlorination roasting systems, CaCl2·2H2O as the sole additive, and an equimolar mixture of CaCl2·2H2O-NaCl, followed by aqueous leaching. The chlorination roasting parameters, including temperature, reaction time, and additive (CaCl2·2H2O) dosage, were investigated. Maximum leachability of Rb 95%, Cs 93%, and K 87% was attained at the following conditions: CaCl2·2H2O 662 kg/t, temperature 900 °C, and reaction time 1.5 h. The equimolar mixture system demonstrated comparable leachability to the CaCl2·2H2O system, presenting a potentially lower-cost roasting alternative, though it increases the sodium load in the subsequent leachate. Thermodynamic analysis using Gibbs free-energy calculations and equilibrium composition diagrams was carried out to identify the reaction mechanisms and phase stabilities, thereby providing a fundamental basis for the observed roasting behavior. This work presents a systematic investigation of the recovery of Rb, Cs, and K from microcline, leucite, and nepheline phases in an SRWL-derived pegmatite residue, providing a thermodynamically grounded framework for multi-metal extraction from complex aluminosilicate matrices.
To address the issues of lacking quantitative correlation between pore structure and rare earth leaching behavior, as well as the unclear classification criteria for pore size intervals during the magnesium sulfate leaching of ion-adsorption rare earth ores, laboratory column leaching experiments were conducted under different void ratios (0.9, 0.95, 1.0, 1.05, and 1.1). Combined with pore characterization techniques such as micro-CT, mercury intrusion porosimetry (MIP), and the centrifuge method, as well as COMSOL Multiphysics numerical simulations, a classification standard for pore size intervals was proposed, and the control effects of different pore size intervals on seepage and ion transport were quantitatively elucidated. The results showed that as the void ratio increased from 0.9 to 1.1, the leaching period shortened by 755.12 h due to pore structure variation, while the peak concentration decreased by 0.40 g/L, indicating that pore structure is the key factor determining the leaching period. Based on cumulative flow rate contribution, pores were classified into three intervals: small (<40 μm), medium (40–85 μm), and large (>85 μm) pores, with medium pores dominating transport channels (74% ~ 78%). When the proportion of large pores increased from 4.19% to 13.25%, the permeability coefficient increased from 5.35 × 10−6 cm/s to 2.85 × 10−5 cm/s. The peak concentration of rare earth ions was the highest and appeared the latest in small pores, while it was the lowest and appeared the earliest in large pores. For every one percentage point increase in the proportion of large pores, the leaching period shortened by approximately 83.4 h on average, and the peak concentration decreased by approximately 0.044 g/L on average. This study provides a reference basis for injection processes and seepage management in ion-adsorption rare earth ores.
In the context of the global energy transition toward low-carbon solutions, nuclear energy's role as a clean source is gaining strategic importance. Given that global reserves of conventional uranium ore are estimated at only 6.3 million tons, the supply of traditional uranium resources is severely constrained. Uranium derived from coal-based byproducts is considered a crucial unconventional resource due to its strategic significance and technological relevance. This study examined the extraction patterns, and mineralogical characteristics of uranium in coal gangue from the Ordos Coalfield. A mechanism for uranium recovery via roasting followed by acid leaching was proposed. Analytical results from SEM, XRD, and XRF showed that uranium microparticles are encapsulated within quartz matrices and kaolinite (the ratio of 7 to 3). Under roasting conditions at 500 °C, clay minerals such as kaolinite in coal gangue undergo a dehydroxylation reaction and transform into amorphous metakaolinite; this process breaks the original aluminosilicate lattice bonds. Although quartz has not yet reached the critical point for the α-β phase transition (573 °C) at this temperature, the physical structural loosening caused by thermal stress and the chemical transformation of the aluminum matrix induce lattice distortion, allowing the uranium components encapsulated within the mineral framework to be fully exposed, thereby significantly enhancing mass transfer efficiency during the acid leaching process. Experiments indicate that the uranium leaching efficiency under these conditions can reach 73.2%; providing a theoretical basis for the efficient recovery of uranium from coal gangue.
The efficient separation of Ta and Sb from Nb in hydrometallurgical processes remains a significant challenge due to their similar chemical behaviors. In this study, a novel peroxo-complexation strategy was introduced to enhance the selectivity of methyl isobutyl ketone (MIBK) extraction in an HF-H2SO4 system. With the addition of H2O2, the extraction efficiencies of Ta and Sb reached 99.6% and 99.7%, respectively, while the co-extraction of Nb was only 2.32%. The selectivity coefficients of Sb/Nb and Ta/Nb significantly increased from 0.63 and 34.7 to 1461 and 6501, respectively. Stepwise stripping using H2SO4 and pure water effectively recovered Nb and Ta/Sb from the loaded organic phase. Raman spectroscopy and quantum chemical calculations elucidated the mechanism of the enhanced separation selectivity. The peroxo-ligand selectively coordinates with Nb to transform it from NbOF52− to NbO2F52−, thereby weakening the negative surface potential of the fluorine atoms bonded to Nb. The binding energy between Nb ions and MIBK decreases from −172 kJ mol−1 to −104 kJ mol−1, and thus the disparity in the affinity of Nb ions and Ta/Sb ions for MIBK is magnified. Overall, this peroxo-coordination approach provides an efficient and simple pathway for the effective purification of Nb solutions.
The co-enrichment of molybdenum (Mo) and tungsten (W) from low-concentration leachates is often constrained by complex anionic matrices, yet the extent to which individual impurity anions interfere with their simultaneous uptake remains poorly understood. Here, a stepwise adsorption study was developed using the strong-base 201 × 7 anion-exchange resin, progressing from a low-concentration synthetic Mo-W reference solution without intentionally added competing anions, to equal charge-equivalent single-anion interference systems, and finally to two real complex leachates dominated by high chloride or high inorganic‑carbon backgrounds. This design allowed the fundamental co-adsorption behavior of MoO42− and WO42− to be distinguished from matrix-induced inhibition. The equilibrium uptake of Mo and W was well described by the Langmuir model within the investigated concentration range, and thermodynamic analysis indicated favorable and endothermic adsorption. Kinetic results revealed a rapid multi-step process involving external mass transfer, transport through the hydrated and swollen polymer network of the gel-type resin, and progressive occupation of exchange sites. Under equal charge-equivalent conditions, the inhibitory strength of the investigated anions followed the order Cl− > SO42− > CO32− > F− at medium and high concentrations, identifying chloride as the strongest competitor. Column experiments showed effective Mo/W co-enrichment from the synthetic reference solution, whereas adsorption performance declined markedly in real leachates because of the combined effects of competing anions and complex solution matrices. The adsorption and characterization results were consistent with an anion-exchange process driven by electrostatic interaction between Mo/W oxyanions and positively charged quaternary ammonium groups. This work provides a quantitative framework for understanding Mo/W co-adsorption from simple to complex low-concentration systems and identifies impurity-anion competition as a critical factor governing the practical performance of 201 × 7 resin.
The effective recovery of zinc and other valuable metals from high-Pb zinc sulfide concentrates remains challenging owing to lead jarosite (Pb-J) formation during oxygen pressure leaching. This hazardous waste byproduct hinders metal extraction and significantly increases the yield of leach residues. This study employs a two-stage oxygen pressure leaching process (LHPL) with a “low-temperature reduction-high-temperature oxidation” strategy to selectively regulate iron leaching and suppress Pb-J phase formation. Compared with the two-stage high-pressure high-temperature leaching process (HHPL), the LHPL process increases the leaching efficiencies of zinc, copper, and indium from 96.7%, 65.2%, and 68.9% to 98.6%, 88.3%, and 90.1%, silver enrichment improved from 72.0% to 93.1%, while residue yield decreases by 20%, thereby effectively minimizing residue generation at the source. Mineralogical analysis reveals that the LHPL process stably converts lead into PbSO4 rather than Pb-J. The LHPL residue exhibits a coarser particle size and hydrophobic surface, leading to excellent settling performance and efficient solid–liquid separation. Moreover, the LHPL process maintains Fe2+ in the first stage to facilitate iron removal and retains soluble Fe3+ in the second stage to accelerate leaching. This approach suppresses Pb-J formation and maximizes iron utilization. These findings provide technical guidance for reducing hazardous iron-containing residues.
The sustainable extraction of manganese (Mn), a critical element used in many high-tech industries, has attracted considerable attention. However, the complex mineral composition makes the processing of manganese oxide ores challenging. This has led to an interest in green chemistry and alternative solvents to reduce environmental impacts. Deep Eutectic Solvents (DES) are recognized for their unique characteristics and environmental benefits. In this study, various DES comprising choline chloride as a hydrogen-bond acceptor (HBA) and organic acids such as oxalic acid (OX), tartaric acid (TA), and lactic acid (LA) as hydrogen-bond donors (HBDs) were synthesized. The key parameters, including HBA:HBD molar ratios, particle size, agitation speed, time, pulp density, and temperature, were systematically studied to optimize manganese recovery. By optimizing the parameters, 99.5% manganese extraction was achieved under the conditions of 40 g/L solid pulp, a molar ratio of 1:2, stirring speed of 200 rpm, temperature of 50 °C, and a leaching time of 1 h using ChCl:OX. Moreover, Density functional theory (DFT) was applied to evaluate the stability of various geometry-optimized Mn2+ complexes with the existing ligands such as OX, LA, TA, and Cl ion, by calculating the Gibbs free energy (ΔG(complex)) for each complex. The theoretical outcomes showed agreement with experimental results. DFT assessment revealed that Mn(OX)₂ has the most negative ΔG(complex) among the DES containing organic acid, followed by tartaric and lactic acid systems, indicating its potential for forming stable complexes with Mn2+. These findings highlight the potential of DES in efficient Mn recovery and simplify the traditional two-step pyro-hydro process by combining it into one step.
The coexistence of scandium (Sc) and titanium (Ti) in titanium dioxide waste acid impedes sustainable Sc recovery due to their analogous physicochemical behaviors in acidic solution. In this study, a ternary synergistic extraction system of Di-(2-ethylhexyl) phosphoric acid (P204), 2-Ethylhexyl phosphonic acid mono-2-ethylhexyl ester (P507) and Tributyl phosphate (TBP) is employed to selectively separate Sc and Ti from the titanium dioxide waste acid. By systematically optimizing key parameters (including pH, extractant concentration, phase ratio, and extraction time), a maximum Sc/Ti separation factor of 167 is achieved. The synergistic extraction system demonstrated exceptional Sc selectivity, achieving 96.8% Sc extraction efficiency with only 15.5% co-extraction of Ti. Slope method, infrared spectroscopy, proton nuclear magnetic resonance/phosphorus nuclear magnetic resonance, and density functional theory calculations, are used to elucidate the mechanism in depth. Mechanistic studies reveal that Sc(III) is efficiently extracted by forming a 1:1:1 mixed cation exchange complex; while Ti(IV), is difficult to effectively dissociate and can only be co-extracted through an inefficient ion association/solvation path. These findings provide a theoretical foundation for optimizing Sc/Ti separation processes in complex titanium dioxide waste acid and offer novel insights into extraction mechanisms, which are of great significance for the green recycling and utilization of strategic rare metal resources.
Separation of Sc from other rare earth metals is a complex and practically important task. In this work three N,O-hybrid ligands, pyridine-2,6-dicarboxamides, 1,10-phenanthroline-2,9-dicarboxamides and 2,2′-dipyridyl-6,6′-dicarboxamides were synthesized and used in polar diluents, dichloroethane (DCE) and meta-nitrobenzotriftoride (F-3), to investigate the extraction and separation of Sc from other REEs. It was shown that for 1,10-phenanthroline-2,9-dicarboxamides, the selectivity of Sc/Ln separation increases from La to Lu. Diamides of 2,2′-dipyridyl-6,6′-dicarboxamides held the highest selectivity for Sc among the three ligands. The selectivity of separation strongly depends on the concentration of nitric acid. The composition of Sc complexes in polar diluent and in single-crystals was identified. The slope analysis showed that the main species in the organic phase is complex with ligand/Sc ratio of 2:1, whereas, XRD-analysis identified 1:1 complexes in the case of single-crystals. Quantum chemical calculations of Sc, Y and Nd complexes of various ligands in various polar diluents were carried out. A comparison of the calculated energies of the complexation reactions showed that the formation of complexes with Sc over other REE is most beneficial.
During zinc electrowinning, strontium carbonate (SrCO3) guides orderly zinc deposition and removes lead impurities. After electrowinning, Sr is enriched in anode slime and crystallized sediments, making its recovery crucial for cost savings. However, the high Ca/Sr ratio and inherent difficulty in separating alkaline earth metals pose significant challenges. In this work, Sr recovery and SrCO3 product preparation from the zinc electrowinning slimes via an integrated process of acid leaching, solvent extraction, stripping and carbonation precipitation was investigated. The Ca removal mechanism from CaSr leachate by P204 solvent extraction was investigated via the FT-IR spectroscopy, equilibrium slope method, and quantum chemical simulations. The research found that saponified P204 exhibits enhanced extraction capacity, but its selectivity diminishes with the increasing saponification degree. At 5% saponification, 33% extractant concentration, O/A of 1, 30 min, pH of 3, and 25 °C, Ca and Sr extraction efficiencies reached 82.7% and 8.55%, respectively. FT-IR revealed that saponification breaks dimeric hydrogen bonds and forms mixed aqueous-organic microemulsion, with extraction occurring via cation exchange without new functional groups. Quantum simulations indicated that the sodium salt from saponification increases the reactivity of hydroxyl oxygen, thereby improving its extraction capability. Furthermore, the energy level difference between CaA2 and the extractant molecules is lower than that for SrA2, leading to a stronger binding affinity for Ca ions. Ultimately, two-stage cross current extraction (O/A = 1/2) with 5% saponified P204 can achieve 99.9% of Ca removal and 16.9% of Sr loss. SrCO3 product with 98.4% purity was obtained via the sodium carbonate precipitation of raffinate.
The accelerating global demand for rare earth elements (REEs), combined with supply concentration and sustainability constraints, has intensified the search for alternative resources. Coal fly ash (CFA), produced in large quantities by coal-fired power plants, represents a promising unconventional REE source. However, the REE potential of coal-derived wastes in Bangladesh remains largely uncharacterized. This study evaluates the occurrence and enrichment of REEs in coal mine tailings and CFA from the Barapukuria coal basin. Twenty composite coal tailing samples and two CFA samples underwent acid digestion followed by ICP-MS quantification of LaLu and Y. Bulk geochemistry was determined by XRF, and mineral associations were examined using SEM-EDS. Coal tailings contain total REE (TREE) concentrations of 8.85–33.96 mg/kg (mean: 21.39 mg/kg) and exhibit consistent LREE dominance (∼60–65%). In contrast, CFA shows marked enrichment, with TREE values of 201.20–243.84 mg/kg (mean: 222.52 mg/kg). UCC normalized patterns display parallel trends between coal tailings and CFA, indicating combustion-driven concentration without significant fractionation. XRF data confirm REE retention within stable aluminosilicate and phosphate matrices during thermal transformation. Elevated levels of Nd, Pr, Dy, and Y in CFA underscore its potential as a secondary REE resource. These findings establish a quantitative foundation for techno-economic evaluation of REE recovery from coal combustion residues in Bangladesh and inform sustainable management strategies for CFA valorization.
