A kind of similar rock material was prepared with gypsum and quartz sand, and influence of seven factors on the strength index of the material was analyzed by performing laboratory tests. The results show that curing temperature and material density are in positive correlation with material strength index, while curing humidity, sand-to-binder ratio, ash-to-paste ratio, fine-grained quartz sand content, as well as water content are in negative correlation with material strength index. Based on the calculation of the sensitivity indices of different influencing factors, it is found that material strength has a sensitivity of only 4.5% to curing temperature, a sensitivity of 104.9% to sand-binder ratio. Factors of curing temperature, water content, material density, curing humidity, ash-to-paste ratio, fine-grained quartz sand content and sand-to-binder ratio are in an ascending order in terms of their sensitivity to uniaxial compressive strength; while factors of sand-to-binder ratio, ash-to-paste ratio, material density, curing humidity, water content, fine-grained quartz sand content and curing temperature are in a descending order in terms of their sensitivity to tensile strength.

An experiment on flotation of calcite in pulp with pH of 8.0 and the dosages for sodium silicate and sodium oleate respectively at 300 mg/L and 1.5 × 10^-4 mol/L was performed by using ultrasonic wave with the frequency at 40 kHz, for investigating the influence of ultrasonic power and sonication duration on the flotation behavior. The variation of sodium oleate content on the surface of calcite was analyzed by adsorption measurement, the chemical status of elements on the calcite surface was analyzed by XPS, and the variation of adsorbed reagents was also analyzed by zeta potential measurement. After 15 min-ultrasonication (40 kHz, 150 W), the flotation recovery of calcite can be increased from 8.5% to 82.7%. It is found that ultrasonication can bring in an increase in the adsorbed amount of sodium oleate on calcite, a decrease in the intensity of Si2p on calcite, and also an increase in the content of C1s. Meanwhile, the deconvolution of O1s spectrum shows reduction in the peak area of sodium silicate, but an increase in that of sodium oleate, and the surface zeta potential of calcite almost unchanged. The results indicate that the ultrasonication can induce a desorption of sodium silicate from the depressed calcite in the pulp containing both sodium silicate and sodium oleate, thus the freshly vacancy can be adsorbed by sodium oleate, leading to the significant improvement in floatability of calcite.

A medium-low grade calcium-magnesium phosphate ore was ground to a fineness of 56.74% -0.075 mm and was used to study the liberation characteristics of particle size within varied ranges on its flotation behavior. The particles were classified into coarse ranges (+0.15 mm, -0.15+0.106 mm and -0.106+0.075 mm) and fine ranges (-0.075+0.038 mm and -0.038 mm). The content of different minerals in each size range was estimated by using X-ray fluorescence spectroscopy and X-ray powder diffraction analysis, while the intergrowth and monomer liberation was evaluated with the backscattering images obtained from scanning electron microscopy. It is found that the content of liberated minerals increases as the particle size decreases. Fluorapatite is primarily enriched in the coarser particles within the size range of +0.075 mm, whereas dolomite is concentrated mainly in fine particles of -0.075 mm. By improving the selective aggregation of valuable minerals and gangues in size range of -0.038 mm, the P₂O₅ loss can be reduced. Furthermore, by reducing the content of coarse particles within the size of +0.075 mm and increasing the content of particles within -0.075+0.038 mm range, both P₂O₅ grade and recovery of the concentrates can be effectively improved.

In processing a mud-bearing high-silver lead-zinc sulfide ore, collectors of dimethylphenyl dithiophosphate and aniline aerofloat were used in combination to collect lead minerals in a low alkali environment, while the noble metal of silver was comprehensively recovered. Furthermore, lime combined with 8372CN were used to depress pyrite in zinc flotation process, thus alleviating the influence of sludge on the quality of zinc concentrate. The closed-circuit flotation test yielded a lead concentrate grading 67.18% Pb at 94.57% recovery, 2 560.37 g/t Ag at 80.54% recovery and 3.60% Zn, and a zinc concentrate grading 51.63% Zn at 93.27% recovery. With this flowsheet, an efficient comprehensive utilization of a mud-bearing high-silver lead-zinc sulfide ore can be actualized.

A process consisting of pre-discarding with heavy medium separation and ambient-temperature flotation was adopted for a scheelite ore sampled from the middle-layer mining zone of a tungsten mine in Hunan Province. A heavy medium separation test with the ore size of -15+0.5 mm as the feeding show that in the heavy medium with the density of 2.37 g/cm³, 31.37% of yielded tailings can be discarded, while the obtained heavy product has WO₃ grade and recovery increased to 0.34% and 96.62% respectively. A test by using a process of desulfurization flotation plus ambient temperature flotation for scheelite demonstrates that the process of pre-discarding with heavy medium has little effect on the operation loss rate of WO₃ in sulfur-containing products. However, the raw ore is directly taken into the flotation without a pre-discarding will lead to more tungsten minerals left in the middlings. The closed-circuit test shows that after a pre-discarding by heavy medium separation, the tungsten concentrate from ambient-temperature flotation has its WO₃ grade and recovery reaching 5.50% and 84.15% respectively. It is concluded that pre-discarding by using heavy medium cyclone is conducive to improvement of tungsten product index and comprehensive utilization of tailings.

A mineral liberation analysis (MLA) automatic measurement technology was used to not only analyze the composition and dissemination of minerals in a skarn copper ore, but also investigate the characteristics like particle size distribution and mineral liberation degree of raw ore and bulk concentrate samples ground to different fineness. After a preliminarily study on beneficiation technique for such ore based on the obtained results, a processing flowsheet comprised sequentially of grinding, Cu-Pb bulk flotation, regrinding of roughing concentrate and Cu/Pb flotation separation. It is shown that such flowsheet with the primary grinding fineness of -74 μm 70% and the regrinding fineness of -20 μm 75% can produce a copper concentrate grading 20.88% Cu at 70.42% recovery, and 183.9 g/t Ag at 76.78% recovery, as well as a product of sulfur concentrate grading 32.65% S at 91.47% recovery.

Rhodococcus opacus (R opacus) was taken as a microbe collector in a flotation experiment to investigate the recoveries of cassiterite, calcite and quartz pure minerals. The results show that, with pulp pH as 4 and R opacus dosage as 1.24 g/L, the recoveries of these three minerals were correspondingly 82.07%, 45.42% and 13.66%. The interaction mechanism between R opacus and cassiterite was investigated by measuring zeta potential and surface tension. Finally, an open-circuit flotation experiment of tin ore was carried out with R opacus as the collector and sodium silicate as the depressant, showing that the grade of tin concentrate was improved from 3.98% to 23.83%.

A study on the separation of a copper-lead-zinc sulfide ore from Yunnan was presented by adopting a Cu-Pb-Zn full selective flotation process. The mineral liberation degree was enhanced by adopting fine grinding of raw ore and selective regrinding of lead roughing concentrate. A combined usage of depressants sodium sulfite and zinc sulfate benefited the synergistic effect and selectivity in depressing, and collectors Z-200, ethionitrogen and BK906 exhibited high selectivity. Under optimized parameters, a copper concentrate with Cu grade of 22.78%, Cu recovery of 83.28%, Pb content of 3.01% and Zn content of 4.23%, a lead concentrate with Pb grade of 75.86%, Pb recovery of 82.75%, Cu content of 0.17% and Zn content of 1.64%, and a zinc concentrate with Zn grade of 51.87%, Zn recovery of 93.16%, Cu content of 0.24% and Pb content of 0.31%, were obtained.

An experimental study was carried out for sulfur reduction and upgrading of a high-sulfur antimony-gold bulk concentrate from Russia with Sb grade of 24.58%, Au content of 76.92 g/t and S content of 14.46%. A closed-circuit flotation test was conducted using a new collector CJ-201 and a new depressant CJ-5S. The Sb and Au grades in the antimony concentrate attained respectively 42.26% and 92.36 g/t, with the corresponding recoveries of 88.04% and 62.18%. With this approach, a antimony-gold concentrate product with high content of antimony and low content of sulfur can be separated out from antimony-gold bulk concentrates collected in a Russian mine.

In order to comprehensively recover fluorite resources from the tailings left after hot flotation of scheelite, a beneficiation process consisting of regrinding, water displacement and concentration, and fluorite flotation was proposed. It is shown that with the feed ore with CaF₂ grade of 25.47%, a fluorite concentrate can be obtained with CaF₂ grade and recovery at 92.35% and 60.59% respectively.