Thermodynamics and Kinetics of Sulfuric Acid Leaching Transformation of Rare Earth Fluoride Molten Salt Electrolysis Slag.

Recovery of rare earths, lithium and fluorine from rare earth molten salt electrolytic slag via fluoride sulfate conversion and mineral phase reconstruction

<p class="1Body">The sulfuric acid leaching of titanium from titanium-bearing electric furnace slag (TEFS) was investigated under different experimental conditions. In the sulfuric acid leaching process, the M<sub>x</sub>Ti<sub>3-x</sub>O<sub>5</sub>(0≤x≤2) and diopside could react with sulfuric acid. The optimum conditions of sulfuric acid leaching process were particle size at < 0.045mm, sulfuric acid concentration at 90 wt.%, acid/slag mass ratio at 1.6:1, feeding temperature at 120 °C, reaction temperature at 220 °C, reaction time at 120minute, curing at 200°C for 120 minute. The [TiO<sub>2</sub>] concentration of the water leaching was 150 g/L, and leaching temperature at 60℃for 120 minute. Ti extraction could reach 84.29 %. F of titanium-bearing solution was 2.15, and the Ti<sup>3+</sup>/TiO<sub>2</sub> of the titanium-bearing solution was 0.068. The TiO<sub>2</sub> content of the leaching residue was 18.32 wt.%. The main mineral phases of the leaching residue were calcium sulphate, spinel, diopside and little M<sub>x</sub>Ti<sub>3-x</sub>O<sub>5</sub>.</p>

Kinetics of rare earth leaching from roasted ore of bastnaesite with sulfuric acid

Sulfuric acid leaching process was applied to extract nickel from roasting-dissolving residue of a spent catalyst, the effect of different parameters on nickel extraction was investigated by leaching experiments, and the leaching kinetics of nickel was analyzed. The experimental results indicate that the effects of particle size and sulfuric acid concentration on the nickel extraction are remarkable; the effect of reaction temperature is mild; while the effect of stirring speed in the range of 400–1 200 r/min is negligible. Decreasing particle size or increasing sulfuric acid concentration and reaction temperature, the nickel extraction efficiency is improved. 93.5% of nickel in residue is extracted under suitable leaching conditions, including particle size (0.074–0.100) mm, sulfuric acid concentration 30% (mass fraction), temperature 80 °C, reaction time 180 min, mass ratio of liquid to solid 10 and stirring speed 800 r/min. The leaching kinetics analyses shows that the reaction rate of leaching process is controlled by diffusion through the product layer, and the calculated activation energy of 15.8 kJ/mol is characteristic for a diffusion controlled process.

Lithium is an important non-renewable resource, and the study of the tiered separation of rare earths and lithium from rare earth molten salt slag is an important approach to solving the current global resource shortage. This article adopts a sulfuric acid leaching process and a lithium-containing solution iron lithium extraction separation process to recover lithium from rare earth molten salt slag. The method systematically investigates the impact of sulfuric acid concentration, liquid-to-solid ratio, leaching temperature, leaching time, pH, P507 concentration, phase ratio, extraction temperature, and extraction time on the lithium extraction effect and iron lithium separation effect in rare earth molten salt slag. The results show that under the conditions of a sulfuric acid concentration of 0.9 mol/L, a liquid-to-solid ratio of 8:1 mL/g, a leaching temperature of 70 °C, a leaching time of 90 min, a pH of 0.9, a P507 concentration of 50%, a phase ratio of 5:3, an extraction temperature of 30 °C, and an extraction time of 20 min, the lithium leaching rate reaches 97.8%, and the separation of iron and lithium is fully achieved. This method can efficiently recover the valuable element lithium from rare earth molten salt slag, which is of great significance for the subsequent preparation of lithium products and the realization of a closed-loop production of rare earth alloys by molten salt electrolysis.

Comparison and evaluation of vanadium extraction from the calcification roasted vanadium slag with carbonation leaching and sulfuric acid leaching

Gossan discarded from mining processes result in metal resource wastage, and its long-term stacking causes environmental hazards. Therefore, this article considers zinc-containing gossan as the research object. The ore was roasted to prepare primary zinc ferrite products and sulfuric acid leaching was performed for purification. Then, XRD analysis was performed to characterize the purified products. The results indicated that the effect of sulfuric acid concentration on the purification of the products was related to its zinc ferrite content. Furthermore, the effect of leaching temperature on the purification of zinc ferrite products was related to sulfuric acid concentration; the lower the sulfuric acid concentration, the more considerable the effect of leaching temperature. The conditions suitable for purifying the products through sulfuric acid leaching are as follows: sulfuric acid concentration of 140 g/L, liquid–solid ratio of 4:1, leaching temperature of 80 °C, leaching time of 120 min, and stirring speed of 300 rpm. This article determines the factors affecting the purification of zinc ferrite by sulfuric acid leaching along with the optimal purification conditions. The findings presented herein provide a theoretical foundation for the development of new processes for preparing zinc ferrite, which has considerable industrial application value.

A novel process for recovering iron, niobium and scandium from Bayan Obo tailings has been developed. In this paper, the treatment of Bayan Obo tailings by Ca(OH)2-coal roasting and sulfuric acid leaching process was investigated. In the Ca(OH)2-coal roasting process, niobium-bearing minerals are converted to CaTiO3 structure of Ca(Ti0.8,Fe0.1,Nb0.1)O3 and Ca2Nb2O7 which are soluble in sulfuric acid. The pyroxene and amphibole that are Sc-bearing silicates are mainly decomposed and reduced into metallic iron which can be recycled by weak magnetic separation. Scandium in the silicates is converted to Sc2O3. In the sulfuric acid leaching process, Ca(Ti0.8,Fe0.1,Nb0.1)O3 and Ca2Nb2O7 are converted to Nb(OH)5 that could easily dissolve in sulfuric acid by dissociating into Nb(OH)4+ and OH− when Sc2O3 is dissolved into heat sulfuric acid. Bayan Obo tailings were roasted with Ca(OH)2-coal at elevated temperature, followed by magnetic separation and sulfuric acid leaching. The optimized experimental parameters are proposed as follows: Ca(OH)2-coal-tailings mass ratio of 20: 5: 100; roasting at 1200 °C for 2 h; the magnetic field (magnetic separation) of 270 mT; the liquid-solid ratio of 4:1 (ml/g); leaching at 245 °C for 1 h. Iron concentrate with a grade of 88.39% and a recovery rate of 91.92% is obtained. The leaching results show that the leaching rates of niobium and scandium could achieve 95.52% and 95.75%, respectively.

The comprehensive recycling of aluminum electrolysis cell waste barrier material is urgent. This study focuses on the sulfuric acid leaching of waste barrier material, systematically examining the effects of factors such as reaction temperature, liquid-to-solid ratio, sulfuric acid concentration, and reaction time on the leaching of elements like lithium, aluminum, sodium, and silicon. The experimental results show that under the conditions of 0.9 mol/L sulfuric acid concentration, a liquid-to-solid ratio of 20:1, a reaction temperature of 90 °C, and a reaction time of 1.5 h, the leaching rates were 84.5% for lithium, 85.6% for aluminum, 98.5% for sodium, and 4.8% for silicon. The sulfuric acid leaching process of the waste barrier material follows a shrinking core model and is controlled by internal diffusion. The apparent activation energies for the leaching reactions of lithium, aluminum, and sodium were 4.29 kJ/mol, 8.99 kJ/mol, and 9.11 kJ/mol, respectively. The selective leaching of lithium, sodium, and aluminum from silicon was successfully achieved in the sulfuric acid leaching of the waste barrier material.

In this paper, the extraction of vanadium from stone coal by roasting with MgO and leaching with sulfuric acid has been investigated, and the mechanism analysis of stone coal roasting with MgO was studied. The results indicated that under the conditions that the mass fraction of the particles with grain size of 0–0.074 mm in raw ore was 75%, the roasting temperature was 500 °C, the roasting time was 1 h, MgO addition was 3 wt %, the sulfuric acid concentration was 20 vol %, the liquid-to-solid ratio was 1.5 mL/g, the leaching temperature was 95 °C, and leaching time was 2 h, resulting in a vanadium leaching efficiency of 86.63%, which increased by 7.73% compared with that of blank roasting. The mechanism analysis showed that the degree of calcite decomposition was low and, thus, magnesium vanadate was more easily formed than calcium vanadate below 500 °C. Moreover, magnesium vanadate was easier to dissolve than calcium vanadate during the sulfuric acid leaching process. Thus, the vanadium leaching efficiency was enhanced by using MgO as a roasting additive below 500 °C. Additionally, at high temperature the formation of tremolite would consume calcium oxide produced from the decomposition of calcite, thus, the formation of calcium vanadate was hindered, and V2O5 would react with MgO to form magnesium vanadate. Therefore, the vanadium leaching efficiency of roasting with MgO was higher than that of blank roasting at high temperature.

Introduction Neodymium magnets contain large amounts of neodymium, praseodymium, and dysprosium. Dysprosium, which is more expensive than neodymium and praseodymium, is added to neodymium magnets to increase their heat resistance. Neodymium magnets are used in numerous consumer electronics and electric- and hybrid-powered vehicles. Large amounts of neodymium are wasted each year when products that contain these magnets are discarded. Processes for the recovery of rare earth elements from neodymium magnets have been developed. Rare earth oxides are recovered from these processes, and these oxides are subsequently reduced to rare earth metals using molten salt electrolysis. We have reported a recovery process for rare earth elements from neodymium magnets using molten salt electrolysis, where these elements were recovered as alloys. In this study, we focus on the electrical behavior of rare earth elements in neodymium magnets during molten salt electrolysis. We use anodic polarization experiments to determine the optimal electric leaching conditions and observe the leaching behavior of rare earth elements. Materials and Methods The electrolysis potential was controlled using a potentiostat. The reactor was made of Pyrex glass and purged with Ar gas. A eutectic salt mixture of 59-mol% LiCl and 41-mol% KCl (melting point: 626 K), which melted at 723 K, was used in the electrolysis bath. The cathode electrode was a glassy carbon rod. The anode electrode was a neodymium magnet, a neodymium rod, a dysprosium rod, or an iron wire. The reference electrode was Ag/AgCl (0.1 N) in a eutectic composition of LiCl–KCl; this electrode was placed in a mullite tube. The electrolysis bath was maintained at 473 K for 24 h under a vacuum to eliminate water. Results and Discussion The scan rate of anodic polarization was 5 mV s−1. The oxidation current of neodymium and dysprosium is generated at approximately −2.2 V, and oxidation potential is the lowest among the elements in the neodymium magnet. The iron oxidation potential was the highest at −0.7 V. Iron is the main component of the neodymium magnet, and the leaching of iron must be avoided in the recovery process. The use of potentiostatic electrolysis enables the selective leaching of rare earth elements from the neodymium magnet, as reported in our previous research. Rare earth elements were leached from the neodymium magnets using potentiostatic electrolysis. The neodymium magnets were used as the anode, and a carbon rod served as the cathode. The electrolysis potential was −1.0 V, and the quantity of electricity was approximately 1200 C. the residual magnet and molten salt compositions were similar for all three neodymium magnets used in the experiments. The residual iron content increased, and the concentrations of the rare earth elements decreased. The rare earth elements were leached from the neodymium magnet into the molten salt, and the total rare earth content in the molten salt was greater than 99.0 mass%. Conclusion Rare earth elements were leached from neodymium magnets using electrolysis in a molten eutectic mixture of LiCl and KCl. The oxidation potential of all neodymium magnet was −1.0 V. The oxidation potential of dysprosium was similar to those for neodymium and praseodymium. The content of rare earth elements in the leaching component was greater than 99.0 mass%. Acknowledgement This work was supported by Environment Research and Technology Development Fund of the Ministry of the Environment, Japan 3K143005.

A novel process for recovery of scandium, rare earth and niobium from Bayan Obo tailings: NaCl-Ca(OH)2-coal roasting and acid leaching

Utilizing Pakistan chromite as raw material, the rapid leaching of chromium and iron could be realized by the sulfuric acid leaching process on the condition of atmospheric pressure and the addition of oxidant A. And the leaching rate of chromium and iron would be 98.5% and 71.9%, respectively. The sulfuric acid leaching processes with different temperature were systematically studied by chemical analysis and phase analysis. The results showed that, with the increase of reaction temperature, the leaching rate of chromium would increase gradually, but the leaching rate of iron increased at first and then decreases and reached its maximum at 140°C. When the temperature > 160°C, the phases of the leaching residue were magnesium iron silicate and a few of silica, no chromohercynite, chrompicotite and magnesioferrite existed in the chromite. The leaching solution of sulfuric acid leaching process could be used for preparing the basic chrome sulfate, and there is no Cr6+ in the leaching residue and solution. The results would provide theoretical guidance for solving environmental pollution problem of Cr6+ in traditional chromate production process.

Natural graphite is classified as a strategic critical raw material by the European Union due to its strategic significance. The demand is steadily increasing, but the current purification methods are not environmentally sustainable, highlighting the need for greener alternatives. This study investigates the effects – sulfuric acid (H2SO4) leaching, thermal treatment, and their combination – on natural graphite from a flotation pilot plant. In the sulfuric acid leaching, the influences of reaction time (60–300 min), temperature (70–100°C), sulfuric acid concentration (0.5–3 mol/L), and liquid-to-solid ratio (10–20 mL/g) were systematically studied using experimental design. Leaching effectively removed 91.4% iron and 52.9% aluminum, but was ineffective against silicon-containing phases. A short 15 min thermal treatment at 2400°C in an argon atmosphere using induction annealing eliminated most silicon phases and other impurities, although residual 2.65 mg/g iron and 1.27 mg/g silicon remained. The combined approach reduced iron and silicon contents to 0.24 and 0.25 mg/g, respectively, and increased the carbon content from 78.4 to 97.6 wt% to near commercial battery-grade levels (98.0 wt%). Additionally, the combination-treated graphite exhibited the lowest degree of structural defects, offering a more sustainable route for purifying natural graphite to high-purity levels compared to conventional halogen-containing techniques.

Study on mechanisms of different sulfuric acid leaching technologies of chromite