Hydrometallurgy最新文献

In-situ leaching is essential for mining rare earth elements (REEs) from ion-adsorption rare earth (RE) ores. The efficiency of RE mining is dependent on the permeabilities of ion-adsorption RE ores, and the permeabilities are controlled by their pore structures. However, the current understanding of the pore structures and permeabilities of the RE ores during leaching is limited, particularly their dynamic variations, the controlling roles of different pore structure parameters on the permeability, the roles of the influencing factors and the mechanisms causing the variations. To investigate the above-mentioned issues, we conducted an experimental study of simulated in-situ leaching on undisturbed ion-adsorption RE ore samples under different conditions via constant waterhead permeability tests. The leaching conditions included leaching time, concentration and waterhead of the (NH₄)₂SO₄ solution. Three-dimensional (3-D) pore structures of the RE ores before, during, and after leaching were constructed via X-ray computed tomography, and their pore structure parameters were measured using 3-D image computation. It was observed that the dynamic variations in both the pore structure parameters and permeability coefficients of the RE ores occurred in three distinct stages: rapid reduction, less rapid reduction, and little reduction. The experimental results also revealed that the permeability coefficients of the RE ores were primarily dependent on the average coordination number among the seven pore structure parameters examined, and that the pore throats larger than 30 μm in diameter served as the most effective seepage channels within the RE ores. The waterhead of the leaching solution had a stronger influence on the variations in pore structures and permeabilities compared to that of the concentration. Analysis of the particle/aggregate size distribution and mineralogical compositions of the RE ores before and after leaching indicated that the decreases in the pore structure parameters and permeabilities were largely attributed to clogging of the pore throats and pores by migrated and newly formed clays, as well as swelling of the latter. The newly formed clays were the products of the decomposition of K-feldspar and mica resulting from chemical reactions with the leaching solution. The clays from the two sources occurred in disaggregated and aggregated forms. The results of this study provide an important reference for the mining of RE ores via leaching.

Cyanidation has long served as the industrial standard in the gold industry, necessitating heightened attention to cyanide management owing to its inherent toxicity. This study developed a novel MnO₂-assisted thiocyanate leaching system as an alternative to cyanide. The sample was generated from a refractory gold concentrate processed by roasting, containing 54.4 g/t gold with iron oxide and quartz as the major mineral phases. Within the framework of the novel system, gold recovery was investigated in terms of gold association, recovery behavior, thermodynamics, and optimization using cyclic extraction protocols. The most suitable leaching conditions were determined as 1.20 mol/L NaSCN, 4.00 mmol/L MnO₂, pH at 1.00, liquid-to-solid ratio of 4 mL/g, and leaching duration of 24 h, resulting in a maximum gold extraction of 94.8%. Subsequent zinc precipitation facilitated a gold recovery of 99.4% from the leachate. Furthermore, the implementation of a cyclic extraction scheme, coupled with reagent addition to offset consumption, yielded a significant thiocyanate dosage reduction by 75% while maintaining gold recovery. This acidic leaching system yields satisfactory gold recovery with minimized reagent consumption, particularly suited for refractory gold ore that has undergone acidic pretreatment.

Given the limitations of current processing methods and the potential of high recovery of silver as a flotation concentrate, this study proposes a pre-oxidation followed by oxygen pressure sulphuric acid leaching route for zinc, copper and indium and thoroughly examines the alteration in micro-morphology of the silver concentrate of 1.22% Pb, 3010 g/t Ag and 649 g/t In. Under the optimal conditions of a sulphuric acid pre-oxidation temperature of 60 °C, and an acid-to-ore mass ratio of 0.6, over 15 h, the PO and -OH polar groups of flotation reagents on the surface of silver concentrate particles are effectively neutralized through sulphuric acid oxidation. This action prevents the formation of a hydration film and mineral inclusions, thereby facilitating the efficient leaching of valuable elements. With a set reaction temperature of 155 °C, an initial acidity of 160 g/L, a reaction time of 180 min, an oxygen partial pressure of 0.8 MPa and a liquid-to-solid ratio of 7 mL/g, the leaching efficiency of Zn, Cu and In in the silver concentrate reach 94.8%, 98.4% and 91.1%, respectively while the lead and silver grades in the leach residue increase to 2.88% and 7754 g/t, respectively.

Cesium is an important strategic resource, and solvent extraction is the most frequently used technology for its extraction from various minerals and brines. However, this common method faces operational and cost problems owing to large alkali consumption. A synergistic extraction system consisting of 4-tert-butyl-2-(α-methylbenzyl) phenol (t-BAMBP) and di-(2-ethylhexyl) phosphoric acid (D2EHPA) was designed and used to extract Cs⁺ from an alkali-free solution. The effects of variables such as: (i) pH, (ii) concentration of t-BAMBP and D2EHPA, (iii) temperature, (iv) Cs⁺ concentration, and (v) coexisting cations, on the Cs⁺ extraction performance of the synergistic system were investigated. The most suitable organic phase composition was 1.0 mol/L t-BAMBP and 0.1 mol/L D2EHPA in kerosene, and the synergistic coefficient was up to 57 at 25 °C and an organic/aqueous phase volume ratio of 1/1. The extraction sequence of cations using the t-BAMBP–D2EHPA synergistic system followed the descending order Cs⁺ > Rb⁺ > Ca²⁺ > K⁺ > Li⁺ > Mg²⁺ > Na⁺. The chemical formula of the extracted species was determined as [CsA(HA)(ROH)₂] using the slope method. The Cs⁺ extraction process is an exothermic reaction with an enthalpy (ΔH^o) of −55.3 kJ mol⁻¹, confirmed by the thermodynamic study. After three-stage countercurrent extraction, the extraction efficiency of Cs⁺ was 92.6%, demonstrating excellent selectivity from coexisting cations. The t-BAMBP–D2EHPA synergistic system showed outstanding economic and environmental advantages and a good application prospect to develop a conceptual process flowsheet for extraction with this system and stripping with HCl to separate cesium salt from salt-lake brine.

Alpha titanium bis(hydrogenphosphate) monohydrate, α-Ti(HPO₄)₂·H₂O, (α-TiP) is a precursor material utilized to obtain a broad range of important compounds having a great deal of applications ranging from ion-exchange chromatography, chemical catalysis to energy storage materials. The novel synthetic procedure developed in this study shows a higher conversion of titanium in ilmenite to α-TiP with a good purity. For the synthesis of TiPs, the direct use of natural ilmenite ore with commonly available chemicals such as KOH, H₃PO₄, HCl, and H₂C₂O₄ highlights the significance of the present investigation. A mixture of ilmenite and KOH with a molar ratio of 1:4 was roasted at 800${}^{°}C$ for 4 h to concentrate all the titanium to potassium titanate and iron to iron oxide. A reaction between potassium titanate and iron oxide with 85% (w/w) H₃PO₄ acid results in a leachate rich in iron in the form of soluble iron phosphates and a white precipitate identified as α-TiP. Calcination of α-TiP at 800${}^{°}C$ produces titanium pyrophosphate, TiP₂O₇. Residual iron that co-precipitated with α-TiP was further removed by multiple washing steps with complexing acids; H₃PO₄ or concentrated HCl. Washing with oxalic acid (H₂C₂O₄) produces a precipitate upon calcination identified as titanyl pyrophosphate, (TiO)₂P₂O₇. Flowcharts were developed and the chemical identities and the purity of the prepared α-Ti(HPO₄)₂·H₂O and (TiO)₂P₂O₇ were tested with X-ray diffraction, X-ray fluorescence, thermogravimetric, and Raman spectroscopic techniques.

In this study, a comprehensive hydrometallurgical processing of molybdenite (MoS₂) concentrate was investigated. This investigation involved a novel approach combining mechanical activation, leaching, and polyelectrolyte extraction methods. The integrated method effectively addressed the challenge of low leaching rate of molybdenite, resulting in the successful production of high-purity molybdenum trioxide. Milled molybdenite samples were analyzed by different methods of X-ray diffraction, atomic force and electron microscopy, and BET. The increasing trend of the specific surface area during milling was determined by a model fitting which was useful for optimization of milling time. Several leaching reagents were studied to achieve high molybdenum dissolution. The most promising results were achieved through a two-hour process, yielding an impressive leaching efficiency of 80% and a resulting Mo concentration of 6700 mg/L. Molybdenum recovery was efficiently carried out through polyelectrolyte extraction, as confirmed by ICP and CHNS analyses, demonstrating selective precipitation of molybdenum from the solution. The subsequent calcination of the precipitated molybdenum(VI) compound resulted in the production of high-purity molybdenum trioxide. Furthermore, a conceptual hydrometallurgical treatment process for molybdenite concentrate was proposed, aiming to recover molybdenum, sulfuric acid, and copper. This proposed process presents a promising avenue for further exploration in pilot plant studies within the molybdenum industry.

Separation and recovery of copper (Cu) from sulphate mediated CuCr spent catalyst leach solution through solvent extraction approach has been systematically investigated. A number of Ionic Liquid (IL) and other conventional extractants were used while investigating the selectivity and efficient extraction tendency of either IL or extractants towards Cu(II). In context of copper extraction efficiency, the adopted organic reagents followed the descending order: R₄PSCN > R₄PD > R₄PCy > Cyphos 101 > Aliquot 336 > D2EHPA > Cyanex 272. The extraction behavior of Cu(II) was established based on slope analysis method. From the results it was noticed that extraction occurs through a cation exchange mechanism with association of a mole of Cu(II) per mole of R₄PSCN. The plot of log D vs. log [R₄PSCN] yield a linear relationship with a slope close to 4 which is used to propose extraction reaction mechanism/ stoichiometry. The nature of complex between Cu(II) and IL was further examined using FTIR analysis of the loaded organic phase with the diluent. Mc-Cabe Thiele diagram was constructed to predict quantitative extraction of Cu(II) which revealed the need for two stages at aqueous to organic (A:O) phase volume ratio of =3:1. The stripping isotherm constructed at optimum NH₄OH concentration (0.4 M) suggests the need for two stages at O:A phase volume ratio of = 2:1 for complete stripping of copper with regeneration of R₄SCN for further use. Both isotherm conditions were validated by 6 cycles of counter current simulation (CCS) study for obtaining the desired amount of Cu(II) loaded R₄SCN (during extraction) and/or stripped copper solution (during stripping). Overall copper enrichment was ∼6 fold leading to produce a copper(II) solution of 48 g/L from 6 g/L Cu(II). The stripped solution was subjected to crystallization study to produce copper sulphate crystals of high purity and was confirmed by XRD analysis of crystal phases.

Efficient separation of lithium from alkaline solution with high Na/Li ratio is of great importance for the development of batteries for new energy industry. This work proposes the use of a novel extraction system composed of ethylhexyl salicylate (ES) and trialkylphosphine oxide (TRPO). The equilibrium experiment revealed that the order of metal ions extracted by ES/TRPO extraction system is Mg²⁺ > Ca²⁺ > Li⁺ > Na⁺ > K⁺. A recovery process including extraction, scrubbing, and stripping was designed to recover lithium from a solution of 3.25 g/L Li, 52 g/L Na and 0.8 mol/L OH⁻. More than 99% of lithium could be extracted to the organic phase. After stripping, a lithium-rich solution containing 25 g/L Li was obtained, and the system showed good stability in the cycling experiments. The FT-IR analysis and DFT calculation were conducted to investigate the extraction mechanism. The results demonstrated that ES mainly coordinates with metal ions through Ph-O and CO bonds to form a hexatomic ring complex, thereby allowing metal ions to enter the organic phase. The calculated binding energies of the complex are highly consistent with the equilibrium experiment. The present work may provide a novel extraction system to efficiently recover lithium from alkaline solution with high Na/Li ratio.

Electronic waste, or e-waste, represents one of the rapidly expanding categories of waste worldwide. By 2019, the global production of e-waste had surged to 53.6 million tons. Due to its substantial metal content, e-waste holds significant financial value, estimated at US$57 billion globally in 2019, predominantly concentrated in printed circuit boards (PCBs). Previous studies have explored hydrometallurgy techniques to extract base metals from PCBs, but effectively recovering these solubilised metals remained a challenge. This research sought to assess metal recovery from PCB waste leachate by utilising hydrogen sulfide generated through a consortium of sulfate-reducing bacteria (SRB) in a fluidised bed reactor (FBR). Both lactate and glycerol were examined as potential organic electron donors for the sulfate reduction. With lactate (1 g L⁻¹) as the electron donor, the FBR achieved an average sulfate reduction efficiency of 62%, with a hydrogen sulfide (H₂S) production rate of 250 mg H₂S-S L⁻¹ d⁻¹ and H₂S-S concentration of 300 mg L⁻¹ in the effluent. When glycerol was the organic electron donor, the average sulfate reduction efficiency was 49%, H₂S production rate was 210 mg H₂S-S L⁻¹ d⁻¹ and H₂S-S concentration was 260 mg L⁻¹. Desulfovibrio, Desulfococcus and Desulfomicrobium were the dominant sulfate reducers in the FBR. The resulting dissolved hydrogen sulfide was employed to recover metals from e-waste leach liquor. Utilising biogenic sulfide and NaOH, a notably high precipitation efficiency (>99%) was attained for aluminum, nickel, copper, and zinc. Additionally, iron, utilised in the e-waste leaching process, was also recovered with an efficiency exceeding 99%. The precipitation of metals occurred within a pH range from 1.5 to 8.5. Overall, this process facilitated the formation of valuable mixed-metal precipitates from waste PCB-derived leachate. These precipitates could undergo further purification or serve as raw material for subsequent processes.

Bayan Obo mixed rare earth (RE) concentrate is one of the most significant rare earth mineral resources worldwide. However, the concentrated sulfuric acid roasting method, which is commonly used to treat ores, generates pollutants such as waste gas, wastewater, and leach residue, resulting in the squandering of the associated resources. This paper introduces a green process involving selective mineral phase transformation (MPT) by heating, followed by leaching to separate bastnaesite and monazite in a mixed RE concentrate to facilitate their subsequent decomposition or extraction. The effects of the MPT conditions on rare earth extraction were investigated. During the MPT process, bastnasite decomposed into REOF, which is more easily leached, whereas monazite remained unchanged. Under suitable conditions, the leaching efficiency of REOF reached 93.7%, while that of monazite (REPO₄) was only 3.2%. Furthermore, the content of monazite in the leach residue was 91.2%. Compared to the mixed RE concentrate, numerous cracks and pores were generated on the surface and inside the MPT products. Furthermore, the total pore volume and specific surface area significantly increased, which made the reaction between the MPT products and hydrochloric acid more efficient. Thus, the MPT process facilitated the leaching of bastnasite and its separation from monazite.