To address problems of cumbersome procedures, time-consuming and high cost for traditional flotation test of molybdenum-bismuth ores, a Bayesian optimization (BO)-based reagent optimization model and a reagent simulation model with combining functions were established for molybdenum-bismuth ore flotation. The simulation functions of flotation reagent were processed with the BO-based model, and it is shown that in a test with 100 sets of computer simulation, the maximum of the function by BO after 25 iterations was 0.93 on average, compared to the optimum value of 0.86 after 25 trials by conventional single-factor test, presenting a 8.14% rise. This verifies the feasible application of BO applied in reagent optimization for molybdenum-bismuth flotation. The laboratory mineral processing data from a concentrator were taken in a study with BO. A model of ore grade and recovery was established via Gaussian process for model fitting. For obtaining the maximum of objective function by multiple iterations, it required an average of 18 iterations by using BO, while 30 trials on average were required by conventional single-factor test. It is concluded that BO can improve the optimization efficiency by 40%.

With sodium metabisulfite as a reducing agent, the copper-cobalt oxide ore of chrysocolla was treated by reductive leaching process in sulfuric acid solution system, and the controlling factors in the process were systematically optimized by adopting a central composite design for response surface methodology. Based on investigation of effects of various factors on cobalt leaching rate, appropriated leaching conditions were obtained for reducing the interference caused by silica gel during the leaching process. The results show that under the following optimal conditions, including sulfuric acid with an initial concentration of 2.0 mol/L, temperature of 80 ℃, liquid and solid in a ratio of 4 mL/g, stirring at a speed of 350 r/min and leaching time of 30 min, the cobalt leaching rate can reach 85.16%.

Slags produced after reduction smelting of Bayan Obo niobium concentrate were taken for research, and laws of influence for basicity on slag viscosity, turning point temperature and viscous-flow activation energy were explored by testing viscosity characteristics of high-temperature smelting slags. Samples taken at the turning point temperature and as the viscosity reaching 8 Pa·s were quenched respectively, and then taken for observing crystalline phases. It is found that the basicity of the slags exerts a significant effect on its high-temperature viscosity characteristics. An increase in temperature leads to a decrease in viscosity, while as basicity increases, the viscosity initially decreases and then increases. Specifically, at a basicity of 1.0, the viscosity attains the minimum value of 0.2 Pa·s at 1 500 ℃; while at 2.0, the viscosity increases remarkably. Both the turning point temperature and viscous flow activation energy at a basicity of 1.0 reach their minimum values, 1 325 ℃ and 243.61 kJ/mol. At different basicity, slags initially have different crystalline phases. At a basicity of 0.5, calcium silicate is the main precipitated phase; at 1.0, perovskite and titanite are precipitated; while at 1.5 and 2.0, perovskite is the dominant precipitated phase.

A mixture of two acids (H₂SO₄, HCl) was used to enhance the acid leaching of vanadium (V) and gallium (Ga) from red mud, and the effects of key parameters on the leaching rates of V and Ga were investigated, including volume ratio of two acids, acid concentration, reaction time, leaching temperature and liquid-to-solid ratio. The results show that under the following conditions, including two acids in a volume ratio of 1∶1, acid concentration of 28.89%, liquid-to-solid ratio of 9, reaction time of 60 min and leaching temperature of 120 ℃, the leaching rates of V and Ga can reach 90.65% and 97.30%, respectively. X-ray diffraction (XRD) and thermodynamic analyses indicate that muscovite, hematite and perovskite in red mud are dissolved to some extent during the acid leaching process, and vanadium and gallium exist in different valence states in the leachate. It is concluded that acid leaching can effectively promote dissolution of minerals and release of valuable metals.

With the dust from a tin smelter as raw material, the occurrence state of tin (Sn) and indium (In) therein was firstly studied. Then, a combined processing technique consisting of water leaching and oxidative acid leaching was adopted to separate and recover the key elements of Sn, In, Zn and Cd from the tin smelting dust, and effects of factors including suluric acid concentration, liquid-solid ratio, reaction temperature and reaction time on the leaching process were explored. The results show that Sn and In in the tin smelting dust mainly exist in the forms of SnO₂, In₂S₃ and Zn₂SnO₄. After water leaching treatment, the dust was leached with sulfuric acid at a concentration of 200 g/L by adding potassium permanganate at a dosage of 1.5%, with liquid-solid ratio of 5.0, at a temperature of 65 ℃ for 120 min, resulting in the leaching rates of In, Zn and Cd at 82.00%, 95.60% and 94.30% respectively, while the leaching rate of Sn at 0.84%. After such treatment, the Sn content in the dust increased from 42.44% to 60.87% with an enrichment ratio of 1.43. The particles of the acid leaching residue were porous on the surface, showing that the dust with dense structure changed to the residue with loose one. This tin-rich acid leaching residue with low impurity content can be back to tin smelting process, and indium can go to the subsequent process for extraction and recovery. By this way, efficient recovery and utilization of valuable metals from tin smelting dust can be actualized.

As for the acid leaching residue of lateritic nickel ore with a TFe grade of 51.38% and S content of 2.01%, a comparative experimental study was carried out for different phase transformation processes. The results show that tests with magnetizing roasting and direct reduction roasting processes generate roasted products containing a wide variety of minerals with complex dissemination characteristics, which is not conducive to subsequent iron-sulfur separation and sorting operations. The lateritic nickel ore, after having been roasted at a hydrogen-rich environment at 900 ℃ for 60 min, is subjected to grinding and magnetic separation, producing a low-sulfur iron concentrate with a TFe grade of 72.50% and a S content of 0.067%. It is concluded that such roasting process at a hydrogen-rich environment can make iron and sulfur effectively separated from each other, which is conducive to subsequent smelting operations.

Iron phosphate dihydrate (FePO₄·2H₂O) is usually synthesized via three technical routes, including ammonium process, sodium process and iron process. A systematical evaluation of its thermodynamic properties is presented, which can reveal the phase transition behavior of iron phosphate dihydrate during the dehydration process, and provide a theoretical basis for selection of technical route for industrial synthesis. The enthalpy change (ΔH), entropy change (ΔS), Gibbs free energy change (ΔG), and logarithm of the reaction equilibrium constant (lgK）of the core reaction of those three processes are calculated by using HSC Chemistry 6.0 software, and the thermodynamic data of dehydration process are verified by performing a thermogravimetry-differential scanning calorimetry (TG-DSC) experiment. The results show that the synthesis by sodium process has the lowest ΔG and the highest lgK, indicating its thermodynamic spontaneity and reaction rate are optimum, followed by the ammonium process, while the iron process presents the worst performance. The TG-DSC experiment results show that the dehydration process of iron phosphate dihydrate is exothermic at 709.9 ℃, corresponding to the transformation of FePO₄ from α-quartz phase to β-quartz phase. Featuring high spontaneity and rapid reaction, the sodium process is considered as a industrial synthesis with thermodynamic advantage.

Spent lithium iron phosphate batteries were taken in a research on lithium extraction by selective leaching process and preparation of high-purity lithium phosphate. It is shown that by adopting a process consisting of leaching with sulfuric acid and hydrogen peroxide, neutralization for impurity removal, and precipitation with sodium phosphate, the lithium leaching rate can reach 98.07% and the iron leaching rate is just 0.28%, after leaching at room temperature with sulfuric acid at a concentration of 0.3 mol/L, hydrogen peroxide at a dosage of 0.50 mL/g, and liquid-solid ratio of 8 mL/g, showing that separation of lithium and iron can be actualized by selective leaching. Then, after the leachate is neutralized with sodium hydroxide for removing impurities, and subjected to precipitation process at a temperature of 90 ℃ for 60 min, with an initial pH of 10.5, sodium phosphate and lithium in a molar ratio of 1∶1, the lithium precipitation rate reaches 96.64%, and a high-purity lithium phosphate product with a lithium content of 16.16% is prepared. It is concluded that by adopting such process, lithium can be efficiently recovered from spent lithium iron phosphate batteries.

In order to investigate the microstructure and mineral composition of sintered limonite (42% in mass fraction) within different size fraction, a sintering pot test was used to simulate the actual sintering process conditions. Thereby, the sintered ore with different particle sizes were prepared, and then mineralogical microscope was used to analyze its microstructure characteristics. The results show that if the mixed ore has the powdered ore with size larger than 5 mm in a mass fraction of 25.23%, the finally sintered ore with small particle size increases, presenting a poor pelletizing performance. As the sintered ore has its particle size reduced from 40 mm to a range of 5-12 mm, the bonding phase therein is in a mass fraction decreasing from 59.10% to 23.19%, and the calcium ferrite has its morphology converting from being acicular to flaky and agglomerated, and its mass fraction decreases from 39.43% to 17.26%. It is found that the sintered ore has converted from the structure with magnetite interwoven with calcium ferrite to granular structure composed of hematite, magnetite and bonding phase. The euhedral to subhedral magnetite grains are transformed to be anhedral, and the number of micropores between and within the grains increases, presenting a "porous and thin-walled" structure, resulting in worse performance in the strength of the sintered ore.

Column leaching experiments were carried out for the high binding-rate copper oxide ore from Mulashi hydrometallurgical plant in Zambia to investigate the effects of proportion of powdered ore (-1 mm), acid consumption for curing and consumption of dripped acid on copper leaching rate. The results show that those factors, such as proportion of powder ore, acid consumption for curing and consumption of dripped acid are in descending order of influence on copper leaching rate. Under the following conditions, including powdered ore accounting for 53%, dripped acid consumption of 20 kg/t and acid consumption of 10 kg/t for curing, the copper leaching rate can reach 72.86%, up by 2.56 percentage points compared to on-site production. Phase and morphology analysis of leaching residue shows that Cu in free copper oxide is mainly leached out by column leaching with dripped acid, while the copper minerals left in residue are bound to copper oxide and copper sulfide. Besides, CaSO₄ precipitate is prone to be formed to cover mineral surface during column leaching, thus affecting the copper leaching rate.