The mechanism for calcite depression with esterified starch was investigated by performing flotation test, surface tension test, zeta potential measurement and infrared spectroscopy analysis. The results show that, compared with ordinary starches, the esterified starch serves as a significantly stronger depressant for calcite. The esterified starch with bipolar functional groups can be chemisorbed onto calcite, reducing the contact angle of calcite surface and increasing its hydrophilicity. It also causes a negative shift of zeta potential of calcite, therewith, enhancing dispersion of calcite particles and effectively reducing hydrous entrainments in flotation products. In the flotation system using sodium oleate as the collector, esterified starch, possessing greater steric hindrance than sodium oleate, can prevail in their competitive adsorption and hinder the adsorption of sodium oleate.

The electronic structure and spin state of pristine rhodochrosite and rhodochrosites with Mn atoms substituted by Ca, Mg or Fe atoms were calculated by density functional theory (DFT), and the magnetism of rhodochrosites was investigated by calculating their spin magnetic moment. The results show that substituting Mn atoms in rhodochrosite for Ca, Mg or Fe atoms can weaken the magnetism of rhodochrosite, and, among the four types of atoms, Mn atoms exhibit the highest saturation magnetization, followed by Fe atoms. The magnetic field and spin magnetic moment generated by the relative motion of ion core to electron of Mn²⁺ are larger than those of Ca²⁺, Mg²⁺ and Fe²⁺.

A complete set of mineral processing technology was proposed for beneficiation tests of a beach placer with TiO₂ grade of 3.33% and Cr₂O₃ content of 0.36%. After a pre-concentration comprised of a gravity separation, a low intensity magnetic separation and a medium intensity magnetic separation, the rough concentrate was processed by roasting followed by dry magnetic separation, producing a titanium concentrate with TiO₂ grade of 47.72%, Cr₂O₃ grade of 0.071% and TiO₂ recovery of 64.76%. It is concluded that the goal to prepare qualified titanium concentrate by extracting titanium and reducing chromium can therewith be achieved.

The process mineralogy and separation feasibility of an ilmenite ore from Xinjiang were comprehensively studied to recover iron and titanium resources from the gravity separation concentrate. The process mineralogy results show that the valuable minerals therein include ilmenite, manganese-bearing ilmenite, titanium-bearing magnetite and magnetite, while the gangue minerals therein are predominantly tremolite, riebeckite, garnet and sphene. After one stage of grinding, iron minerals were collected with one stage of low intensity magnetic separation. Then, two stages of high intensity magnetic separation were adopted to recover titanium minerals. The titanium rough concentrate was treated by a flotation for desulfurization and a flotation process consisting of one stage of roughing, one stage of scavenging and four stages of cleaning. The whole process produced an iron concentrate grading 63.93% TFe at 69.11% recovery, presenting a good separation index, and a titanium concentrate was obtained with grade of 32.59% TiO₂ at 28.04% recovery. It is found that low TiO₂ content and high content of iron-bearing silicate gangue minerals are the main factors affecting the ilmenite beneficiation indicators.

A kind of silver concentrate in Henan Province has the associated lead therein with an average grade around 7%, lower than the grade for pricing. A classification test was conducted with a hydrocyclone for the silver concentrate, and effects of parameters, including cone angle, underflow orifice diameter, and angle-cone ratio, on the classification and concentration of silver concentrate were explored. It is found that by adopting a 100 model hydrocyclone with cone angle of 60°, feed in a mass fraction of 13.89% to be fed at the pressure of 0.2 MPa, overflow pipe of 22 mm in diameter, and the underflow orifice of 12 mm in diameter, a high-lead silver concentrate grading 6316.29 g/t Ag and 12.39% Pb can be obtained at corresponding recoveries of 79.06% Ag and 79.45% Pb, with a yield of 49.25%; meanwhile, a low-lead silver concentrate with silver grade of 1 623.36 g/t can be obtained. This processing technique can obtain both high-and low-lead silver concentrate at the same time, leading to the economic value of the silver concentrate improved by 687.96 Yuan per ton.

A prediction method for mineral processing indices of shaking table was proposed based on eXtreme Gradient Boosting (XGBoost). As for the separation product zone, it can quantify the colour difference by colour moments, the shape characteristics by image moments, and the texture characteristics by evaluation indices of grey-level co-occurrence matrix, including contrast, homogeneity, correlation and ASM energy. The features of the separation product zone can be effectively extracted according to the difference in its colour, shape and texture during separation process. Subsequently, those features can be filtered by employing XGBoost, with which a prediction model can be then constructed. After being trained with the test set to predict the grade, recovery rate and yield of concentrate, this model can achieve accurate prediction of separation indices. It is found that the XGBoost-based model outperforms the decision tree model and random forest model in terms of accuracy when applied for predicting recovery and yield of concentrate.

An experimental study with a novel flotation reagent was carried out for some refractory low-grade polymetallic ore from Henan Province, and a process of Cu-Mo bulk flotation, followed by Cu/Mo separation was adopted in the experiment. With a processing flow consisting of a Cu-Mo bulk flotation (including two stages of roughing, one stage of cleaning, and two stages of scavenging) with L43 as a collector and L56 as a frother, followed by regrinding of the obtained Cu-Mo bulk concentrate, and Cu/Mo separation (including one stage of roughing, three stages of cleaning and three stages of scavenging) with L789 as a depressant, the experiment produced a molybdenum concentrate grading 50.780% Mo, 0.602% Cu with 65.30% Mo recovery and 4.72% Cu recovery, and a copper concentrate grading 0.165% Cu and 0.362% Mo with recoveries of 4.49% Cu and 1.61% Mo. It is shown 90.79% Cu is recovered into the mixed tailings.

To solve problems of poor selectivity, thick foaming and large reagent consumption in recovering spodumene with fatty acid collectors, anionic and cationic collectors were introduced to flotation tests of spodumene, feldspar and quartz. According to the results of pure mineral tests, a anionic-cationic collector can enhance the synergistic effect of reagents, and improve the grade and recoveries of concentrates. CFLH-18, a combination of anionic and cationic collectors, was introduced in a close-circuit flotation test of granite pegmatite-type spodumene from Jinchuan County of Sichuan Province, and a spodumene concentrate with Li₂O grade of 4.62% was collected at 78.54% recovery. Compared with the on-site practice using fatty acid collectors, Li₂O grade and recovery are improved by 0.10 percentage points and 5.93 percentage points. Meanwhile, the concentrate enrichment ratio is raised from 3.77 to 4.57.

A technique consisting of acid leaching, oxidative precipitation for impurity removal, and solvent extraction was adopted to effectively separate and recover nickel from high-nickel electroplating sludge, and then high-purity nickel sulfate hexahydrate was prepared. Firstly, the sludge was leached with sulfuric acid, and most of nickel, copper, iron and aluminum in the sludge were leached out in the leachate; then, hydrogen peroxide was added to the leachate for oxidation and precipitation to remove iron therein, and the suitable endpoint pH value was 3; after that, sodium sulfide with an appropriate excess coefficient of 1.2 was added to remove impurities therein, resulting in the precipitation rates of Fe and Cu reaching 99.43% and 98.13% respectively; by adjusting the pH value of the precipitated solution, nickel hydroxide was precipitated and then dissolved with sulfuric acid into nickel sulfate solution, which was extracted with P204 for deep removal of impurities. As a result, the secondary extraction rates of Cu, Fe and Al were up to 98.41%, 99.99% and 95.15% respectively, and the extraction rate of Ni was just 9.4%. The obtained raffinate was subjected to evaporation concentration, followed by cooling crystallization, and a nickel sulfate hexahydrate with purity exceeding 99.95% was finally prepared.

To reasonably select the composition of molten slag for efficient enrichment of rare earth elements therein, the viscosity and melt structure of mixed rare earth slag under different binary basicity conditions were explored using rotating cylinder method, FTIR, Raman spectroscopy, and X-ray photoelectronic spectroscopy. The results show that with the binary basicity of 0.50, 0.86, 1.00, and 1.20 respectively, the corresponding viscosity of the slag system at 1 550 ℃ is 1.30, 0.87, 0.60, and 0.36 Pa·s, and the activation energy of the viscous flow unit is 268, 220, 216, 202 kJ respectively; the area fractions of Q¹ (Si) and Q² (Si) in the molten slag are all more than 24%. As the binary basicity increases, the area fraction of bridging oxygen (O⁰) decreases, and the area fractions of non-bridging oxygen (O^-) and free oxygen (O^2-) decrease, and the K value representing the polymerization degree of slag gradually decreases from 0.53 to 0.10. It is found that the structure system of molten slag tends to be simple with the increase of binary basicity, which is the main reason for decrease in the viscosity of a mixed rare earth slag.