The tailings from gravity separation of cassiterite in a polymetallic ore from Guangxi Zhuang Autonomous Region were taken for experimental research. Due to its low grade, fine particle size of target minerals therein and similar floatability of tourmaline and cassiterite, a multiple ligand metal-group collector, CSC-BHA-Pb, and a high-efficiency depressant, XY1, were adopted to selectively inhibit gangue minerals dominated by tourmaline while efficiently enrich fine-grained cassiterite. On the basis of iron removal and desulfurization, an open-circuit flotation test of cassiterite, consisting of one roughing, one scavenging and seven cleaning, finally produced a tin concentrate with a Sn grade of 14.05% and a recovery of 28.79%, achieving efficient tin enrichment.

With a copper sulfide ore from Jiangxi Province taken in the research, an experimental study was carried out for enhancing the recovery of associated Au by flash flotation. In a laboratory test, the underflow of secondary-stage classification was subjected to an open-circuit flash flotation consisting of two-stage roughing and two-stage scavenging, leading to the Au grade up from 16.60 g/t to 140.60 g/t, presenting good floatability. Additionally, the Cu grade in the concentrate from flash flotation cell is higher than that in the main process, which proves that flash flotation of classification underflow will not affect the Cu recovery from the main process. Industrial practice demonstrates that with the ores of different Au grade, the Au recovery rate has increased after technical transformation. A size analysis by sieving reveals that those fully liberated minerals within conventional size fraction in classification underflow can be mainly recovered by flash flotation, thus being prevented from overgrinding due to returning to the grinding mill.

The flotation reagent system was optimized for a lead-zinc ore in Qinghai based on experimental studies. It is shown that a new chelating agent (PT-8) combined with the conventional 25^# xanthate can effectively enhance the flotation recovery of copper-lead sulfide minerals. A new collector (SXF-10) at a low dosage can efficiently recover marmatite from the ore, significantly improving the recoveries of Cu, Pb and Zn from flotation process. With raw ore at a grinding fineness of -0.074 mm 70%, 25^# xanthate and PT-8 are used for Pb flotation, and SXF-10 is used for Zn flotation. A closed-circuit test with such process flow can produce a Cu-Pb concentrate grading 3.52% Cu and 62.53% Pb at corresponding recoveries of 70.08% and 93.75%, respectively, and a Zn concentrate grading 48.19% Zn at 87.97% recovery.

A polymetallic sulfide ore in Qinghai assays 1.45% lead and 2.64% zinc, and is associated with copper, sulfur, iron, plus rare and precious metals (gold and silver). Process mineralogy of the ore was studied by multi-elemental analysis, phase analysis, and MLA (Mineral Liberation Analyzer) analysis. A closed-circuit test was performed after condition tests on this basis. In the general mineral processing of lead-zinc sulfide ores, lime is added through the whole flow, which is not conducive to comprehensive utilization and recovery of copper, gold and silver. Aiming at this problem, a processing concept was proposed for comprehensive utilization of the valuable elements therein, which consists of preferential flotation of lead under natural pH value, zinc flotation by inhibiting sulfur from the obtained lead flotation tailings, and finally iron recovery after sulfur removal of the tailings from the previous step. Results show that with the polymetallic sulfide ore ground to a fineness of -0.074 mm 70%, the valuable elements therein can be recovered for comprehensive utilization by adopting a combined flotation and magnetic separation process, consisting of copper-lead flotation, zinc recovery from the obtained copper-lead flotation tailings, sulfur recovery from the obtained zinc flotation tailings, and finally iron separation from the obtained tailings. It is shown that the lead concentrate grading 60.12% at recovery of 94.72%, zinc concentrate grading 46.99% Zn at recovery of 88.26% and iron concentrate grading 67.22% Fe at recovery of 9.33% are produced.

Based on analysis of chemical composition, mineral composition, main mineral properties and disseminated grain size of slag from smelting of nickel laterite ore, a series of exploratory experiments were carried out. The iron in the slag was recovered by magnetizing roasting and magnetic separation. The magnetizing roasting is performed for 40 min at 750 ℃ with a coal powder at a ratio of 2%. The roasted ore is subjected to a two-stage low-intensity magnetic separation (LIMS) after being ground to a fineness of -0.045 mm 80%, and an iron concentrate grading 60.38% Fe can be obtained at recovery of 71.55%.

An experimental study on flotation separation was conducted for a gold-antimony concentrate with grade of Sb and Au at 20.66% and 98.95 g/t, respectively, from Russia. In an experiment on antimony floatation while suppressing sulfur by adding efficient sulfur depressant of BK526, activator of lead nitrate, and collector of ammonium dibutyl dithiophosphate, the gold-antimony concentrate was ground to a fineness of -0.045 mm 86%, sodium sulfide and activated carbon was added to remove reagent and sulfuric acid was used to adjust the pulp pH to be acidic. Finally, it produced an antimony concentrate grading 51.56% Sb at 86.06% recovery, and a gold concentrate grading 108.35 g/t Au and 4.41% Sb, with gold recovery of 71.20%. It is shown that stibnite and gold-bearing pyrite can be efficiently separated.

An experimental study was carried out to address problems of higher grade of iron in the tailings from 1^st-stage high intensity magnetic separation (HIMS) and higher metal losses in the production of Lilou Iron Mine in Anhui Province. In the study, the low-intensity magnetic separator (LIMS) used before 1^st-stage HIMS was replaced with a CCT-A high-gradient permanent magnetic drum separator, and the HIMS process consisting of one roughing and one scavenging was optimized to a HIMS process consisting of one roughing and two scavenging. With these measures, both strongly and weakly magnetic minerals can be effectively recovered. The 1^st-stage scavenging tailings have the iron grade decreased by around 1 percentage point, and nearly 40 000 tons of iron concentrate can be recovered annually, which can bring remarkable economic benefits to the mine.

In Meishan Iron Mine, the iron concentrate produced on the site has SiO₂ grade more than 6.0%. Aiming at this problem, an experimental study was carried out based on the microscopic identification of different concentrate products by using a new process flow, in which the magnetic field intensity of low-intensity magnetic separation (LIMS) was reduced for roughing, and a process of regrinding and re-separation was added for the scavenging concentrate by high-intensity magnetic separation (HIMS). The effects of those technical transformation on the TFe grade of iron concentrate and the content of impurity SiO₂ therein were explored. The results show that after such technical transformation, the LIMS can produce a concentrate grading 63.57% TFe and 3.37% SiO₂ with corresponding recoveries of 72.79% and 12.94%, with a yield of 52.21%; while the HIMS can produce a concentrate grading 44.79% TFe and 9.07% SiO₂ with corresponding recoveries of 17.20% and 11.67%, with a yield of 17.51%. This new flowchart can produce the concentrate grading 58.85% TFe and 4.80% SiO₂ at corresponding recoveries of 89.98% and 24.61%, with a total yield of 69.72%. It is shown the SiO₂ content in the total concentrate of Meishan Mine can be greatly reduced.

Based on flotation tests of a low-grade copper sulfide ore, a processing technique was developed, composed of flash flotation of Cu, Cu-S bulk flotation, and Cu-S separation after regrinding of roughing concentrate from bulk flotation. Effects of factors on flotation were also explored, including grinding fineness, flotation reagent, regrinding fineness of roughing concentrate from bulk flotation. Ammonium dibutyl dithiophosphate was used as the collector for Cu in the flash flotation, a combination of ammonium dibutyl dithiophosphate and butyl xanthate as the collector for Cu-S bulk flotation, lime as a sulfur depressant, and terpineol oil as a frother. With grinding fineness of -0.074 mm 60% and regrinding fineness of -0.045 mm 70%, a closed-circuit test for the raw ore assaying 0.40% Cu and 2.45% S produced a bulk copper concentrate grading 19.49% Cu at 87.68% recovery, and a sulfur concentrate grading 48.65% S at 35.11% recovery. It is concluded that efficient separation and enrichment of copper and comprehensive recovery of sulfur from this low-grade copper ore can be actualized by this technique.

A roasting and acid leaching process was adopted to extract lithium from a low-grade clay-type lithium ore with a Li₂O grade of 0.13%. After exploration of the effects of various factors on lithium leaching rate, including roasting temperature, roasting time, and mass fraction of sulfuric acid, liquid-solid ratio, leaching temperature and leaching time in acid leaching process, the optimal experimental conditions were determined as follows: roasting at 500 ℃ for 1 h, leaching at 95 ℃ for 60 min with sulfuric acid at a mass fraction of 15%, and liquid-solid ratio of 6 mL/g. Under these conditions, the lithium leaching rate can reach 85.26%, indicating that efficient leaching of lithium from such low-grade clay-type lithium ore can be actualized. XRD and SEM analyses of the samples before and after leaching show that lithium leaching is attributed to the exchange between hydrogen ions in the sulfuric acid solution and lithium ions in the mineral.