Based on investigation of the influence mechanisms of acid and alkaline corrosion for flotation separation of lepidolite and feldspar, efficient separation of lepidolite and feldspar in different particle size fractions was achieved by flotation with surface dissolution and collector interaction. It is found that in flotation process, alkaline corrosion can promote the separation of lepidolite and feldspar more effectively than acid corrosion. Under optimal conditions, lepidolite concentrates in size fraction of 38-74 μm and-38 μm are obtained with recoveries of Li₂O at 92.54% and 83.20%, respectively. The effects of acid and alkaline corrosion on the wettability, morphology, reagent adsorption and electrical potential of the mineral surface were explored by contact angle measurements, high-speed camera observations, SEM, AFM, XPS, infrared spectroscopy and zeta potential analyses. Results show that alkaline corrosion, due to its distinct mechanisms at the surfaces of lepidolite and feldspar, can promote the separation between them; whereas acid corrosion has no selectivity for two minerals. Besides, both acid and alkaline corrosion are independent of particle size of minerals.

In consideration of biotite content in iron concentrate from magnetic separation, it was proposed that a reverse flotation process with sodium oleate was adopted, in which soluble starch was used to depress magnetite and ferric ions (Fe³⁺) to activate biotite. And then, the flotation results and the mechanism of reagent interaction in flotation process were verified by performing a pure mineral test, zeta potential measurement, adsorption capacity detection and infrared spectra analysis. It is found that Fe³⁺ can slightly enhance adsorption of sodium oleate and soluble starch onto magnetite surfaces, but significantly enhance their adsorption onto biotite. With Fe³⁺ at concentration of 1 × 10^-3 mol/L, a reverse flotation test of the artificially mixed ore of magnetite and biotite (with magnetite accounting for 70% by mass) can produce biotite at a recovery of 62.76% and magnetite at a recovery of 16.09%, indicating a good separation effect.

For solving the difficulty in separation between zircon and rutile, 1-hydroxyethane-1, 1-diphosphonic acid (GS) was independently synthesized, and effects of GS, etedronate sodium (ETD) and amino-trimethylphosphonic acid (ATMP) on flotation of zircon and rutile were also explored. Based on micro-flotation tests, the collecting performance and corresponding mechanism of collectors were analyzed by contact angle measurement, Zeta potential measurement and FTIR spectroscopy. It is found that GS exhibits excellent collection performance and selectivity in separation under neutral conditions, bringing in significant improvement in separation between rutile and zircon. Phosphonic acid group of GS enhances the electronegativity and hydrophobicity by multi-point adsorption, resulting in effective collection. The FTIR analysis has further approved chemical adsorption of phosphonic acid group.

A 3D finite element simulation model of a ZH high-intensity magnetic separator (HIMS) was established based on Ansys Maxwell. The magnetic field distribution characteristics and the variation law of magnetic field intensity in the separation channel of the separator with three kinds of medium (serrated plate, elliptical rod and cylindrical rod) were systematically analyzed. A simulation experiment on separating particles was conducted by using the discrete element method. The results show that the serrated plate medium produces an obviously focused magnetic field effect at the serrated part, with the maximum field intensity reaching 2.08 T. The periodic strong gradient fluctuation is beneficial to the efficient capture of fine ilmenite particles. The elliptical rod medium can bring good uniformity in magnetic field strength, which is suitable for stable separation of coarse minerals. The cylindrical rod medium leads to a wide magnetic field distribution and can be applicable to separation of mixed materials with wide size fraction. The simulation results are in good agreement with the experimental measurement, which verifies the reliability of the model. Further particle simulation reveals that serrated plate medium can break particle agglomeration and suppress non-magnetic mineral entrainment by periodic gradient reversal, and finally realize the efficient enrichment of ilmenite. It is concluded that the serrated plate medium has significant advantages in separation of fine ilmenite and is the preferred medium for the ZH HIMS.

Medium and low grade phosphate ores were floated in a packed flotation column, and the phosphate recovery after different floatation time were measured, which were then fitted with four classical kinetic models respectively. Results show that the first-order kinetic model is more suitable to describe and predict the flotation process. The particle residence time in the flotation process of phosphate ore measured in the experiment was correlated to the flotation kinetic constant, and a mathematical model was derived for the average residence time and flotation kinetic constant. With this model, the appropriate average particle residence time is calculated to be 201.87 s and the corresponding flotation kinetic constant is 0.114 81.

A lead-zinc sulfide ore with high sulfur content in a mine always causes calcium scaling in pipelines in practical production, and also requires strong acid activation in the subsequent sulfur floatation. In consideration of these disadvantages, a new reagent was proposed to be used in flotation process under alkaline conditions for improving the floatation indicators. Then, an experiment was conducted by adopting a process of preferential flotation of lead-zinc-sulfur in sequence under low-alkaline conditions with a new collector (BN+311) and a new activator (H1). It is shown that with the raw ore assaying 4.20% Pb, 9.39% Zn and 30.27% S, the sequential flotation of lead, zinc and sulfur yields a lead concentrate grading 55.37% Pb (containing 1.98% Zn and 24.95% S), a zinc concentrate grading 53.49% Zn, and a sulfur concentrate grading 48.07% S (containing 0.50% Pb and 0.26% Zn); while the contents of Pb and Zn in the tailings are 0.66% and 0.37%, respectively. By reducing the on-site pulp pH from 12 to about 9, qualified lead and zinc concentrates can be produced by adopting this process, and the sulfur concentrate can also be obtained with low contents of Pb and Zn, presenting excellent flotation indicators.

With iron phosphate slag as raw material, Fe and P elements were extracted by acid leaching, which was then used to synthesize battery-grade iron phosphate. Meanwhile, activated carbon and carbon nanotubes were introduced to adsorb the residual Fe and P elements in the leachate, and then high-performance Fe/P oxygen reduction catalysts (FP-6) were prepared by calcination. The results show that the FP-6 catalyst exhibits a half-wave potential of-0.037 V (vs. Hg/HgO) in 0.1 mol/L KOH solution, demonstrating excellent oxygen reduction catalytic activity. With FP-6 as the cathode catalyst, the zinc-air batteries can deliver a specific capacity of 780 mAh/g and a power density of 121 mW/cm², and stably cycle for 182 h at a high current density of 5 mA/cm². This study provides a low-cost and sustainable approach for efficient recycling of iron phosphate slag and resource utilization.

In order to promote resource recycling from spent lithium-ion batteries, a novel process consisting of acidic roasting and water leaching was proposed to recover valuable metals therein. With LiNi_1/3Co_1/3Mn_1/3O₂ as raw material, effect of additives in roasting process on valuable metals was investigated. The results indicate that obvious weight loss and endothermic behavior occur when LiNi_1/3Co_1/3Mn_1/3O₂ is mixed with additives of NaHSO₄·H₂O, KHSO₄, Na₂S₂O₇, and K₂S₂O₇ at molar ratios of 1∶3, 1∶3, 1∶1.5, and 1∶1.5 respectively and roasted. In the presence of additives, Li, Ni, Co and Mn in the roasted LiNi_1/3Co_1/3Mn_1/3O₂ exist as corresponding sulfates, and the roasted products are massive and compact in structure. After the mixture of LiNi_1/3Co_1/3Mn_1/3O₂ and additives is roasted at 600 ℃ for 0.5 h and then the obtained calcine is leached with water, the leaching rates of metal elements are 99.99%, 98.97%, 99.89% and 98.86% by correspondingly using additives of NaHSO₄·H₂O, KHSO₄, Na₂S₂O₇, K₂S₂O₇, respectively.

To recover molybdenum from copper smelter soot, a process consisting of NaOH leaching, N235 extraction, ammonia stripping and precipitation-roasting was employed. Effects of various factors on molybdenum recovery during leaching and extraction-stripping processes were investigated. The results show that under the following optimal leaching conditions, including NaOH at concentration of 6 mol/L, leaching at 80 ℃ for 1.5 h and a liquid-to-solid ratio of 6∶1, leaching rate of molybdenum is 82.44%; under the following optimal extraction and stripping conditions, including extractant consisting of 30% N235, 10% sec-octanol and 60% diluent, extraction for 5 min with phase ratio (O/A) of 1∶15 and stripping for 6 min with phase ratio (O/A) of 2∶1, the extraction and stripping rates of molybdenum are 99.67% and 92.95%, respectively. Finally, a molybdenum trioxide (MoO₃) product is produced by precipitation and roasting, with the mass fraction of molybdenum upgraded to 60.08% from 0.458% in raw material, presenting an enrichment ratio of 131.18. It is concluded that this processing technique can effectively separate and recover molybdenum resources from soot, providing a reliable technical route for efficient recovery of molybdenum from secondary resources.

To obtain key parameters such as evaporation surface velocity, evaporation surface temperature and flow conditions of liquid metal beryllium during high-temperature vacuum distillation, and then to optimize the temperature field, flow field and equipment accordingly, Ansys Fluent was used to simulate the distributions of temperature and flow fields of liquid metal beryllium in distillation at completion rate of 5%, 20%, 40%, 60%, 80% and 95%, respectively. Four heating modes, including constant-temperature bottom heating, constant-temperature side heating, constant-temperature side-and-bottom heating and constant-power side-and-bottom heating, were compared with each other for investigating advantages and disadvantages. The results show that, compared with constant-temperature bottom heating or constant-temperature side heating, the constant-temperature side-and-bottom heating mode can effectively mitigate severe drop in the overall average temperature in the late distillation stage, which is conducive to reducing energy consumption and maintaining a stable temperature field, and also brings overall uniformity in heating. Based on a comparison, it is found that constant-temperature heating mode is better than constant-power heating in terms of average temperature and temperature uniformity.