Search Results (43)

To address the challenges of low KCl recovery and high collector consumption during flotation at low temperature, a novel approach with utilizing a compound collector consisting of octadecylamine hydrochloride (ODA) and alcohols (butanol, octanol, and dodecanol) to enhance low-temperature KCl flotation recovery was proposed in this study. The flotation performance and underlying mechanisms of this novel amine–alcohol compound collector were investigated through combination of micro-flotation tests, contact angle measurements, and molecular dynamics simulations. The results revealed that KCl flotation recovery decreased with declining temperature using single ODA as the collector, and the maximum KCl flotation recovery was approximately 40% with an ODA concentration of 1 × 10⁻⁵ mol/L at the temperature of 0 °C. Moreover, amine–alcohol compound collector shows different KCl flotation recovery; among them, dodecanol (DOD) presents the best performance at 25 °C with an ODA concentration of 3 × 10⁻⁶ mol/L. The KCl flotation recovery initially increased and then gradually decreased with increasing the DOD concentration, and 90% KCl recovery was achieved with a DOD concentration of 1.5 × 10⁻⁵ mol/L (DOD:ODA = 5:1 in mole) under 25 °C. Furthermore, this compound collector exhibited high selectivity for KCl/NaCl flotation. Mechanism studies indicated that the trend in contact angle changes on the KCl crystal surface closely mirrored the trend in flotation recovery. Molecular dynamics simulations demonstrated that at 0 °C, the presence of DOD resulted in a higher diffusion coefficient for ODA molecules compared to the system without DOD. Additionally, the water molecules in System 3 exhibited a lower diffusion coefficient and a greater number of hydrogen bonds. This novel compound collector offers a potential solution for improving KCl recovery and reducing ODA consumption during low-temperature flotation. It holds significant theoretical and practical implications for advancing low-temperature KCl flotation technology. Full article

The mineral processing of a low-grade tungsten-molybdenum ore (LGTMO) was investigated to assess the potential of recovering molybdenum (Mo) and tungsten (W). Techniques such as Polarizing Microscope (PM), Scanning Electron Microscopy-Energy Dispersive Spectroscopy (SEM-EDS), Mineral Liberation Analysis (MLA), and Advanced Mineral Identification and Characterization System (AMICS) were employed. The recoverable metals in the ore are Mo (0.158% ± 0.03%) and W (0.076% ± 0.02%). Mo exists in two forms: 63.30% as molybdenite and 36.7% as powellite (CaMo_xW_1−xO₄). W is present as 75.26% scheelite and 24.74% powellite. The complete dissociation rates of molybdenite and scheelite-powellite are 27.14% and 88.87%, respectively. Particles of scheelite-powellite with a diameter less than 10 µm account for 34.61%, while molybdenite particles with a diameter below 10 µm make up 72.73%. Scheelite-powellite is mainly associated with olivine and dolomite, while molybdenite is mainly associated with pyroxene, calcite, and hornblende. Based on the process mineralogy, the mineralogical factors influencing the flotation recovery of molybdenite and scheelite-powellite were analyzed. Finally, a complete hydrometallurgical leaching test was carried out. The optimal experimental conditions are as follows: liquid-solid ratio of 6 mL/g, KMnO₄ concentration of 0.015 mol/L, Na₂CO₃ concentration of 0.12 mol/L, NaHCO₃ concentration of 0.024 mol/L, leaching time of 4 h, and leaching temperature of 85 °C. Under these conditions, the leaching efficiencies of Mo and W reach 79.23% and 41.41%, respectively. This study presents a novel approach for the recovery of refractory W and Mo resources in LGTMO while simultaneously providing a theoretical basis for the high-efficiency utilization of these resources. Full article

Lithium, a critical element for clean energy and modern technologies, plays an indispensable role in advancing renewable energy storage, electric vehicles, and high-tech industries. The rapidly growing demand for lithium, along with its limited global production, has led to concerns about the sustainability of current extraction and processing technologies for efficient lithium recovery. This comprehensive review explores global trends in lithium processing, focusing on spodumene beneficiation and extraction techniques. While highlighting well-established conventional processes, such as dense media separation (DMS), flotation, high-temperature roasting, and acid or alkali treatment, it underscores the environmental and economic challenges of these processes, particularly when applied to low-grade lithium ores, which are increasingly being targeted to meet lithium demand. Innovative methods, such as microwave irradiation, are also explored for their potential to improve process efficiency, reduce energy consumption, and minimize environmental impact, offering promising pathways to overcome the limitations of traditional lithium recovery techniques. A significant contribution of this review is its focus on the largely untapped lithium resources of Kazakhstan, presenting geological insights and the potential for sustainable development. By addressing knowledge gaps and integrating technological, eco-friendly, and regional development perspectives, this study provides valuable insights for advancing lithium processing toward more sustainable and circular practices aligned with global climate and resource efficiency goals. Full article

In the wake of global energy transition and the “dual-carbon” goal, the rapid growth of electric vehicles has posed challenges for large-scale lithium-ion battery decommissioning. Retired batteries exhibit dual attributes of strategic resources (cobalt/lithium concentrations several times higher than natural ores) and environmental risks (heavy metal pollution, electrolyte toxicity). This paper systematically reviews pyrometallurgical and hydrometallurgical recovery technologies, identifying bottlenecks: high energy/lithium loss in pyrometallurgy, and corrosion/cost/solvent regeneration issues in hydrometallurgy. To address these, an integrated recycling process is proposed: low-temperature physical separation (liquid nitrogen embrittlement grinding + froth flotation) for cathode–anode separation, mild roasting to convert lithium into water-soluble compounds for efficient metal oxide separation, stepwise alkaline precipitation for high-purity lithium salts, and co-precipitation synthesis of spherical hydroxide precursors followed by segmented sintering to regenerate LiNi_1/3Co_1/3Mn_1/3O₂ cathodes with morphology/electrochemical performance comparable to virgin materials. This low-temperature, precision-controlled methodology effectively addresses the energy-intensive, pollutive, and inefficient limitations inherent in conventional recycling processes. By offering an engineered solution for sustainable large-scale recycling and high-value regeneration of spent ternary lithium ion batteries (LIBs), this approach proves pivotal in advancing circular economy development within the renewable energy sector. Full article

In the face of the global depletion of natural resources and increasing demand for sustainable development, processing industrial waste, such as tailings from processing plants, is becoming essential. This study focuses on combined processing technologies, including flotation concentration and concentrate processing, allowing the efficient recovery of valuable components. This study aims to investigate the possibility of thermochemical enrichment and the opening of low-grade copper tailings of processing plants with the transfer of silicon and rhenium in the form of silicate-ions and perrhenate-ions into a solution with the output of a multifactor multiplicative model and obtaining tabular nomograms. Multifactor experiments on the thermochemical enrichment of rough copper concentrates made it possible to construct partial dependences of silicon and rhenium extraction into a solution and to obtain multiplicative Protodyakonov–Malyshev models of these processes and multifactor nomograms over a wide range of temperatures, durations, and alkali-to-concentrate ratios to determine the maximum recovery rates. The developed multifactor models made it possible to establish the optimal intervals of changes in the concentrate sintering parameters, providing high recovery rates (over 85% of silicon and 98% of rhenium) during subsequent water leaching. Optimal sintering conditions (temperature of 350 °C, the duration of 90 min, and the ratio of NaOH to concentrate = 1:2) ensured a recovery of up to 85% of silicon and 98% of rhenium from the concentrate into the solution. This recovery rate reduces the need for primary raw materials and positively affects the production’s environmental performance because it minimizes the amount of industrial waste disposal. Full article

The effective and environmental separation of chalcopyrite and molybdenite has long presented a challenge in mineral processing due to their similar floatability and close association at room temperature. This study explores the non-toxic 1-thioglycerol (1-TG) as a selective depressant for chalcopyrite–molybdenite flotation separation. An impressive separation effect was realized through single-mineral and mixed-mineral flotation experiments when using 1-TG as a depressant and kerosene as a collector. Contact angle measurements, zeta potential tests, and Fourier transform infrared spectroscopy (FT-IR) confirm the selective adsorption of 1-TG on the chalcopyrite surface, leading to enhanced surface hydrophilicity and the inhibition of collector adsorption. The depression mechanism is further elucidated through X-ray photoelectron spectroscopy (XPS), which demonstrates that it occurs via chemosorption between the thiol group in 1-TG and active iron sites on the chalcopyrite surface. These findings provide a potential efficient depressant for chalcopyrite–molybdenite flotation separation with low dosage, environmental friendliness, and human harmlessness. Full article

Phosphogypsum is an industrial byproduct that is limited in its high-value application due to the presence of dyeing impurities (such as organic matter and carbon black). The flotation method has been verified to be effective in separating these dyeing impurities from gypsum. In this study, microemulsion was used as the collector method of dyeing impurities for their separation from gypsum. The results of flotation tests showed that the microemulsion collector exhibited excellent collection capability and selectivity under natural pH conditions (pH = 1.5). With a microemulsion collector consumption of 400 g/t, purified gypsum of 65.1% whiteness, 95.74% yield, and 97.01% recovery was obtained. The purified gypsum of 65.1% whiteness, 95.74% yield, 97.01 recovery obtained by a used microemulsion collector amount of 400 g/t was better than using the same dosage of kerosene collector. The dispersion behavior of the microemulsion collector was studied by low-temperature transmission electron microscopy. The microemulsion collector demonstrated superior dispersibility, as it forms nano-oil droplets with an average size of 176.83 nm in the pulp, resolving issues associated with poor dispersibility observed in traditional kerosene collectors. Additionally, the nano-oil droplets effectively adsorbed onto the surface of dyeing impurities through hydrogen bonding, enhancing their hydrophobicity. Therefore, the microemulsion collector holds great potential for application in flotation whitening processes involving phosphogypsum. Full article

A new type of amino acid surfactant, lauroyl lysine (LL), is used as a green collector for the low-temperature flotation of quartz. The micro-flotation test results indicate that, under flotation conditions of 10–40 °C, pH = 11.0, 20 mg/L CaCl₂, and 60 mg/L LL, the highest recovery of quartz by LL could reach up to 97.08%. The temperature at which flotation occurs little impacts LL collection efficiency. In contrast, sodium oleate (NaOL) gives inferior performance to LL at all tested temperatures. The adsorption measurement and SEM-EDS results confirm that a quantity of LL is absorbed onto the quartz surface at low temperatures. Investigations into the interaction between the reagents and mineral surfaces are conducted using X-ray photoelectron spectroscopy (XPS) analysis, zeta potential measurements, and Fourier transform infrared (FT-IR) spectra. Findings indicate that LL is adsorbed onto the quartz surface through hydrogen bonds and intense chemisorption. Additionally, the amide groups in the LL molecular structure increase the solubility of the collector at low temperatures, and simultaneously, the amide bond can form an intermolecular hydrogen bond between O and H, which is conducive to quartz flotation. Full article

This study employs microwave roasting to decompose smithsonite mineral (zinc carbonate) into zinc oxide, which then reacts with pyrite to sulfurize its surface, forming zinc sulfide. This process is beneficial for the flotation recovery of zinc oxide minerals. The surface sulfidation behavior of smithsonite under low-temperature microwave roasting conditions is examined through X-ray diffraction (XRD), X-ray photoelectron spectroscopy (XPS), scanning electron microscopy (SEM), and thermodynamic calculations. XRD and thermodynamic analysis indicate that smithsonite completely decomposes into zinc oxide at 400 °C. Introducing a small amount of pyrite as a sulfidizing reagent leads to the formation of sulfides on the surface of decomposed smithsonite. XPS analysis confirms that the sulfide formed on the surface is zinc sulfide. SEM analysis reveals that sulfides are distributed on the surface of smithsonite, and the average sulfur concentration increases with the pyrite dosage. Microwave-assisted sulfurization of smithsonite (ZnCO₃) was found to significantly enhance its floatability compared to conventional sulfurization methods. The optimal mass ratio of ZnCO₃ to FeS₂ is approximately 1:1.5, with the best temperature being 400 °C. These findings provide a technical solution for the application of microwave roasting in the efficient recovery of smithsonite through flotation. Full article

This study focuses on the purification and evaluation of the high-purity quartz (HPQ) potential of vein quartz ore from Pakistan. Vein quartz is grayish-white and translucent, with its mineral composition mainly comprising quartz crystal. Processed quartz sand is obtained from quartz raw ore through purifying technologies, including crushing, ultrasonic desliming, flotation, high-temperature calcination, water quenching, hot pressure acid leaching, and chlorination roasting. The microscopic characteristics show that the vein quartz raw ore has a medium-coarse granular metacrystalline structure, high quartz content, with only a small quantity of fine-grained K-feldspar. The inclusions primarily consist of large-sized primary inclusions and secondary fluid inclusions developed along the micro-fractures, and the content of inclusions in most areas of the crystal is very low or even nonexistent. The quartz ore with such inclusion characteristics is considered a relatively good raw material for quartz. Component analysis shows that the main impurity elements in the quartz ore are Al, K, Ca, Na, Ti, Fe, and Li, with a total impurity element content of 128.86 µg·g⁻¹. After purification, only lattice impurity elements Al, Ti, and Li remain in the processed quartz sand, resulting in a total impurity element content of 24.23 µg·g⁻¹, an impurity removal rate of 81.20%, and the purity of SiO₂ reaching 99.998 wt.%. It is suggested that when the quartz raw ore contains high content of lattice impurity elements, such as Al, Li, and Ti, it is difficult to remove them by the current purification method. In industrial production, considering the economic cost, if quartz sand still contains high content of lattice impurity elements Al, Ti, and Li after flotation, it cannot be used as a raw material for high-end HPQ. Full article

This paper focuses on the environmental leaching of antimony, a critical mineral, using deep eutectic solvents. Mining residues often contain embedded antimony, posing environmental risks. Deep eutectic solvents, known for being low in toxicity, cost-effective, and environmentally friendly, present a promising avenue for sustainable antimony extraction. The study focuses on optimizing the leaching process through experimental analysis by considering variables such as temperature, time, and percentage of solids. Different deep eutectic solvent (DES) compositions are being studied, including choline chloride with malonic acid, thiourea, and ethylene glycol in different molar ratios, to identify the most effective solvent system for antimony extraction. A sample, originally obtained from mining waste produced via the flotation of antimonite ore, was used to test these three types of DESs. By optimizing the leaching process by changing the ratio of solid and liquid components, as well as the amount of oxidizing agent up to 3 g, iodine, yields of up to 100% were achieved after leaching for 4 h at 100 °C. The aim of the study is to advance sustainable resource management by providing knowledge on an ecological and feasible method of extracting antimony from mine waste, leading to more conscious and efficient resource practices in the mining sector. Full article

Zinc extraction from oxide ore has been paid more and more attention to due to the exhaustion of zinc sulfide ore resources. In this work, the volatilization kinetics of Zn from the flotation products of low-grade lead–zinc oxide ore during carbothermal reduction in the temperature range of 900–1300 °C were investigated. The phase transformation in briquettes during the reduction process was investigated by XRD and EPMA. The results showed that the transformation of ZnS by CaO may begin within the temperature range of 900–1050 °C, with the main occurrence observed in the range of 1050–1250 °C. The kinetics behavior of Zn volatilization was associated with the phase transformation. The volatilization of Zn was controlled by the interfacial chemical reaction within 900–1150 °C. As the reaction proceeded, the generation of the product layers (CaS, FeS and new slag phase) impeded the internal diffusion of Zn, CO and CO₂. At this time, internal diffusion served as the rate-controlling step for Zn volatilization in the range of 1150–1300 °C. Hence, a staged kinetics model of Zn volatilization during carbothermal reduction in the form of carbon-bearing briquettes was established, and the apparent rate constants (k(T)) and activation energies (E_a) were obtained. This work provides a scientific basis for the flotation products treatment by carbothermal reduction and is of great significance in improving the sustainability of resources in the zinc smelting industry. Full article

Flotation performance can significantly be affected by seasonal variations due to the changes in water temperature and pulp temperature, type and concentration of dissolved ions in process water. Extreme temperature conditions could be the major factor affecting flotation performance and mask the influence of water chemistry. Therefore, the interactive effects of the temperature and water chemistry should be taken into consideration, particularly for mine sites experiencing extreme temperature conditions. In this paper, effects of temperature, sulphate (SO₄²⁻), thiosulphate (S₂O₃²⁻) and calcium (Ca²⁺) ions on the flotation performance of a Cu-Pb-Zn complex sulfide ore were investigated using a statistical experimental design and modelling approach. The results were evaluated using ANOVA and regression analysis to determine the significant parameters and derive individual regression models for each flotation response using Design Expert software version 6.0.8. Individual regression models were developed for mass pull, water recovery, grade and recovery of the sulfide minerals using the statistically significant main effects and their interactions. The models were used to determine the concentration of the dissolved ions and pulp temperature required to achieve the maximum zinc recovery, maximum zinc grade or the optimum zinc grade and recovery. The results showed that the water chemistry (i.e., the concentrations of Ca²⁺, SO₄²⁻ and S₂O₃²⁻) affected the flotation performance significantly at low temperature (25 °C). At high pulp temperature (60 °C), however, the temperature was the dominant parameter and masked the effects of water chemistry. Details of the statistical experimental design, discussions of the effects of experimental factors and their interactions on flotation performance, and the development of regression models are presented in this paper. Full article

Foam is a medium-stable system composed of gas and liquid phases, which has the advantages of low density at the gas phase and high viscosity at the liquid phase, and has a wide application in oil and gas field development and mineral flotation, but its special medium-stable system also brings many problems in industry applications. Scientists have carried out extensive analyses and research on the foam stability and bubble-bursting mechanism, which initially clarified the rules of bubble breakage caused by environmental factors such as temperature and pressure, but the mechanism of bubble bursting under the action of internal factors such as liquid mineralization and oil concentration of the films is still not clearly defined. In this paper, we propose a compound salt-resistant foaming agent, investigated the influence of the aggregation and adsorption behavior of oil droplets on the liquid films and boundaries, and established a relevant aggregation and adsorption model with the population balance equation. We put forward a liquid film drainage mechanism based on the distribution, aggregation, and transport of oil droplets in the liquid films, so as to explain the changes in foam stability under the action of oil droplets. On the other hand, the viscoelastic analysis of foam fluid is performed with a rheometer, and the results show that in comparison with conventional power-law fluid, foam fluid has a complex rheological behavior for low shear thickening, but high shear thinning. Full article

The resource utilization of cyanide tailings has significant environmental and economic benefits. The efficient recovery of gold from low-grade cyanide tailings containing 1.71 g/t Au was performed by a chlorination roasting–flotation process. The effects of roasting temperature, calcium chloride, internal coke, external coke, copper sulfide concentrate, and kaolin on the recovery rate of concentrate, gold grade, and sorting efficiency were investigated. The optimized process parameters were as follows: 16 wt% calcium chloride dihydrate, 6 wt% internal carbon, 1 wt% external carbon, 9 wt% copper sulfide concentrate, 2 wt% kaolin, and roasting temperature of 730 °C. The sorting rate, gold grade, and recovery rate of gold concentrate can reach 88.48%, 33.46 g/t, and 76.7%, respectively, and the gold grade of tailings was as low as 0.17 g/t. In the matte phase, gold can be enriched in the form of gold grains. Therefore, through chlorination roasting, the trapped gold can be released and deposited on the surface of the matte phase due to the transformation from hematite to magnetite. The gold-deposited metal sulfide can be effectively recycled through flotation. These results have potential guiding significance for the efficient recovery of gold from cyanide tailings. Full article