Based on the process mineralogy study, an experimental study on grinding-separation technique was performed for JISCO's magnetite obtained from its suspension magnetization roasting process. In the experiment, a 110 mT low-intensity magnetic separator was firstly adopted for pre-concentration. The middlings were reground and processed by magnetic separation for enrichment and tailings discarding, and the obtained concentrate was subjected to reverse cationic flotation. Finally, the whole process produced the comprehensive concentrate grading 60.06% TFe and containing 5.17% SiO₂, with metal recovery of 84.27%. In comparison with the previous processing technique, the TFe grade of concentrate is improved by 1.96 percentage points and the metal recovery is increased by 1.38 percentage points, while the SiO₂ content falls by 1.06 percentage points.

An experimental study on a metallurgical grade fluorite concentrate with CaF₂ content of 70.34% was conducted to upgrade fluorite concentrate quality by reducing impurities therein. The results of the full process test show that by adopting a process of high intensity magnetic separation for reducing silicon, followed by flotation of non-magnetic minerals for reducing calcium, the experiment produces a fluorite concentrate grading of 91.95% CaF₂ at 76.52% recovery, with a yield of 58.30%. It is shown that the concentrate meets the raw material standards for anhydrous hydrofluoric acid production.

SAG mill grinding and conventional crushing-grinding processes adopted in Baima Iron Mine are introduced. Based on the detailed comparison of processing flowsheet, production index, equipment operation parameter, construction investment, operation cost, it can be concluded that the SAG milling process has the advantages of simple process, low construction investment, good adaptability to ores and fine grinding fineness, also some disadvantages like high equipment failure rate, short service life for lining plate, relatively low equipment operation rate and high production cost. These results can be referred in the construction of similar mines.

High gradient magnetic separation technique was adopted to separate the (tantalum-niobium)-tin bulk concentrate obtained from gravity separation. The effects of magnetic intensity, pulsating frequency, magnetic rod matrix diameter and rotating speed of ring on (Ta-Ni)/Sn separation efficiency were studied. The test results show that, from the bulk concentrate with grades of Ta, Ni and Sn respectively as 11.65%, 9.33% and 19.59%, the final tantalum-niobium concentrate containing 22.31% Ta, 17.38% Ni and 3.22% Sn can be collected with the recoveries of Ta, Ni and Sn respectively as 95.48%, 92.88% and 8.20%, respectively. Meanwhile, a tin concentrate with Sn grade of 35.87% and Sn recovery of 91.80% can be obtained. The tantalum-niobium and tin minerals can be effectively separated.

The dissolution characteristics and mechanism of heavy metal ions (As, Cr, Cu, Mn, Ni, Pb) and Na⁺, K⁺ in red mud were studied in a simulated natural environment. The results show that the dissolution of heavy metal ions, Na⁺ and K⁺ out of red mud can be promoted by reducing pH value, increasing contact time, raising leaching temperature, and increasing liquid-solid ratio. After dissolution, the characteristic peak values of calcite, magnesium chlorite, gosilicate and triploidite significantly become lower, and elements of As, Cr, Ni, Cu, Mn, Na and K, after dissolution out of red mud, exist in the forms of As₂O₃, Cr³⁺, NiO, CuO, MnO, Na⁺ and K⁺.

Based on the experience from production practices, an orthogonal experiment was performed by simulating a grate-rotary kiln process to investigate factors affecting compressive strength of preheated pellets and also find measures to improve the process. It is found that both OA iron ore and OB iron ore are fine-grained, with -0.074 mm grain content greater than 85%, which meets the basic requirements for particle size in pelletizing process. With 90% OA iron ore blended with 10% OB iron ore, the obtained pellet after drying and preheating process can have FeO content reduced from 26% to less than 4%, with the oxidation degree greater than 85%. With the following optimal thermal parameters of chain grate, including a drying temperature of 200 ℃ for air blowing and a drying temperature of 360 ℃ for ventilation, 720 ℃ for I-stage preheating and 990 ℃ for II-stage preheating, the obtained preheated pellet can have a compressive strength of 1 056 N/pellet. It is verified that the experimental result is in accordance with the anticipation.

Fluoride-containing industrial wastewater was treated by using a composite coagulant of polymerized aluminum chloride (PAC) and polyacrylamide (PAM), and effects of PAC dosage, PAM concentration, pH value, reaction time and stirring speed on fluoride removal efficiency were explored. The results show that by adding 4 400 mg/L PAC, with PAM concentration of 1 mg/L, solution with pH of 7.0, reaction time of 20 min and stirring speed of 300 r/min, the fluoride removal rate can reach 97.98%, leading to the residual concentration of fluoride ion at only 0.81 mg/L. However, the presence of impurity ions can reduce the fluoride removal rate by PAC.

In order to determine a new type of gold leaching system, which can not only bring good leaching effect with low consumption of thiosulfate, but also make the gold in the lixivium to be efficiently recovered, the leaching effects of gold with thiosulfate catalyzed by different combination of metal ions and organic ligands were compared, and then the gold in the lixivium was recovered by resin adsorption. The results indicate that the Ni²⁺-glycine-thiosulfate system can bring good Au leaching effect, showing Au leaching rate up to 89.72% with thiosulfate consumption of 15.28 kg/t. During the gold leaching process in this system, glycine catalyzes the coordination of Au⁺ with S₂O₃^2- and also stabilizes Ni²⁺ under alkaline conditions, and Ni (C₂H₄NO₂)₂⁰ catalyzes the reduction of O₂. When Au is recovered by resin adsorption, Ni ions in the lixivium are hardly adsorbed on the resin, and do not compete with Au ions for adsorption, resulting in Au adsorption rate up to 98.13%. Then, only one-step process is required for desorption of gold-loaded resin, that is, Au can be directly desorbed.

Experimental study was carried out on continuous leaching of mixed wolframite-scheelite concentrate with soda in a horizontal continuous autoclave reactor, and the effects of various processing parameters on the decomposition rate of the mixed ore were investigated. Under the processing conditions, including grinding fineness of -0.050 mm for the mixed ore, leaching temperature of 200 ℃, liquid-to-solid ratio of 2.1∶1, soda usage at four times of the theoretical dosage, addition of NaOH at 25 kg/t of ore, and feeding rate of 10 m³/h, the decomposition rate of tungsten in the mixed wolframite-scheelite concentrate reaches 99.55%, with only WO₃ content of 0.28% in the slag. It is found that the continuous leaching with soda offers the advantages of easy operation, higher production efficiency, lower energy consumption, stable decomposition rate, as well as higher automation.

A two-step acid leaching process was used to leach scandium and yttrium from red mud. Phosphoric acid was used in the first stage of leaching to achieve dealkalization, desilication and enrichment of rare earth elements (scandium and yttrium). Then, the leaching residue was leached with sulfuric acid in the second stage. The effects of sulfuric acid concentration, liquid-solid ratio, reaction temperature and reaction time on the leaching process of scandium and yttrium were investigated, and the leaching kinetics was also analyzed. It is found that leaching of red mud with phosphoric acid can bring better dealkalization and desilication effect, showing the leaching rates of K, Na and Si at 75.62%, 87.82% and 80.25%, respectively. Then, in the second stage of leaching process with sulfuric acid concentration of 2.5 mol/L and liquid-solid ratio of 7∶1, reaction runs at 70 ℃ for 45 min, leading to the average leaching rates of Sc and Y up to 74.49% and 80.02%, respectively. It is shown that the leaching processes of Sc and Y with sulfuric acid follow the liquid-solid reaction model, which is controlled by diffusion, and the apparent reaction activation energies are calculated to be 18.50 kJ/mol and 16.74 kJ/mol, respectively.