Natural chalcopyrite (CuFeS₂) was used as a catalyst to activate persulfate (PS) for butyl xanthate wastewater treatment. The influences of CuFeS₂ dosage, PS dosage and initial pH of solution on the removal rate of butyl xanthate in CuFeS₂/PS system were investigated, and the recyclability of CuFeS₂ was also evaluated. The results show that CuFeS₂ can significantly activate PS. After 10 min reaction with initial pH of 3, CuFeS₂ dosage of 4 g/L and PS dosage of 5 mmol/L, the removal rate of butyl xanthate reached 99.23%, while the removal rate of total organic carbon (TOC) was only 31.29%. After the reaction time was extended to 60 min, the TOC removal rate reached 68.31%, and CuFeS₂ still sustained good catalytic activity after being recycled for five times. The UV-Visible absorption spectra of butyl xanthate wastewater degradation showed that butyl xanthate was initially oxidized to alcohols, then decomposed and mineralized to carbon dioxide and water. X-ray diffraction and fourier transform infra-red spectroscopy analysis showed that CuFeS₂ reacted with PS without surface passivation, thus could be reused. A free radical quenching test and an electron spin resonance analysis showed that the removal mechanism of xanthate included free radical degradation and CuFeS₂ physisorption, in which sulfate radical (SO₄^-·) and hydroxyl radical (HO·) were the main active substances to promote xanthate degradation, and the SO₄^-· played a dominant role.

A kind of ore from Inner Mongolia containing 1.89% Pb, 1.84% Zn and 125 g/t Ag was taken in a flotation experiment for recovering the valuable elements therein. From the process mineralogy study, it is found that the Pb and Zn therein exist mainly in the form of galena and sphalerite, while Ag is in the form of freibergite. A preferential flotation process was adopted for treating this ore, in which zinc sulfate and sodium sulfite were taken as sphalerite depressants, propyl xanthate, ammonium dibutyl dithiophosphate and 25^# dithiophosphate were taken as the collector for lead flotation, and then copper sulfate was taken as the activator and butyl xanthate was taken as the collector for zinc flotation from the tailings of lead flotation. A lead concentrate grading 52.71% Pb, 2.29% Zn, containing 3 182.00 g/t Ag was produced at 94.22% Pb recovery and 82.00% Ag recovery, and a zinc concentrate grading 46.11% Zn, 0.76% Pb, containing 255 g/t Ag was also yielded at 87.31% Zn recovery and 6.78% Ag recovery, showing that silver was mainly pulled into lead concentrate. It is concluded that this processing technique can bring not only a good Pb/Zn separation effect, but also an effective recovery of Ag.

Based on the study of process mineralogical properties and grindability of an iron ore from Australia, experiments were carried out to explore its beneficiability. The results show that this kind of ore has a TFe grade of 34.90%, with magnetite and hematite as the dominant Fe-containing minerals, and quartz as its dominant gangue mineral. The ore is medium hard, with a protodyakonov coefficient of 3.13, and also comparatively hard to mill, with a Bond ball milling work index of 13.27 kW·h/t. On this basis, a flowsheet consisting of sequentially primary grinding, a roughing process of low-intensity magnetic separation (LIMS), secondary grinding and a cleaning process of LIMS was adopted in an experiment, producing an iron concentrate grading 65.45% TFe at 65.98% recovery, with mFe recovery of 95.69%. After intensified grinding, the iron concentrate was subjected to further magnetic separation, resulting in ultrapure iron concentrate grading 71.90% TFe at 60.20% recovery. Meanwhile, the middlings was subjected to a scavenging, and a super-high grade iron concentrate with TFe grade of 69.87% was obtained with recovery of 35.21%, indicating that this ore has a good beneficiability.

An accumulative bauxite ore in Yunnan containing 42.13% Al₂O₃, 15.73% SiO₂ and 17.44% TFe, namely, with an Al/Si ratio of 2.68, was taken in a flotation for Al/Si separation, and the effects of pH regulator dosage, type and dosage of depressant, collector dosage and grinding fineness on desilication performance of flotation were all investigated. A closed-circuit flotation test performed with the optimized parameters produced an aluminum concentrate with Al₂O₃ grade and recovery of 52.16% and 70.30%, Al/Si ratio of 7.69, and TFe grade of 20.89%, respectively. Then, a flowsheet consisting of magnetic roasting and magnetic separation was introduced to treat this aluminum concentrate for Al/Fe separation, which produced a final aluminum concentrate containing 69.34% Al₂O₃ with Al/Si ratio of 8.70, and a rough iron concentrate grading 52.34% TFe with a process recovery of 76.03%.

A lead-zinc deposit is rich in valuable metals of gold and silver. In order to improve the economic benefits of products, the associated gold and silver should be enriched in lead concentrate to the maximum extent. Based on the actual situation in the production, the distribution of associated gold and silver in the products after mineral processing was ascertained, and then optimization tests were perform. The results showed that most of the associated gold and silver were concentrated in the lead concentrate, and the recovery of gold and silver was positively correlated with the recovery of lead. Without altering the original reagent regime for lead flotation, ethyl xanthate was introduced as an auxiliary collector in a close-circuit test adopting a flowsheet consisting of one stage of roughing, two stages of cleaning and two stages of scavenging. And the obtained lead concentrate with grade around 50% can have a high enrichment of gold and silver and a low content of zinc. It is concluded that gold and silver resources can be comprehensively recovered, which can bring in the maximized economic benefit.

The fine slime collected from the flotation overflow in a titanium concentrator located at Panzhihua-Xichang region, with fineness of－0.025 mm 78.10% and TiO₂ grade of 8.33%, was taken as the raw material for study. Based on the characteristics of this slime, a pre-enrichment process was adopted for impurity removal in an experiment, and CC, a new type of dispersant with good selectivity, was used in the following flotation. An ilmenite concentrate was obtained grading 47.97% TiO₂ at 73.09% recovery, indicating the titanium resources in the slime can be reclaimed.

In the actual production, not only the copper oxide ore has a high content of slime, but also the copper minerals therein are prone to be lost during the beneficiation. In consideration of these problems, a cyclone desliming system was added to process the tailings after the copper sulfide was preferentially floated. The industrial experiment results show that a copper oxide concentrate grading 19.20% Cu can be obtained with the process recovery of 76.24% by adopting a direct sulphidizing flotation process to reclaim copper oxide minerals from copper sulfide flotation tailings, and the copper grade of tailings can be reduced to 0.94%. After copper sulfide flotation tailings were deslimed by cyclone, a process consisting of flotation of copper oxide minerals from underflow and a wet leaching to recover copper resource from fine slime (overflow) was adopted, resulting in a comprehensive copper oxide concentrate with Cu grade of 24.63% and process recovery of 85.55%, and the final tailings with Cu grade reduced to 0.60%. The indices of the copper oxide concentrate from the flowsheet containing cyclone desliming process are obviously superior to those by direct sulphidizing flotation process. Besides, this flow is relatively simple with reduced unit consumption of reagent for copper oxide flotation.

The amphibolite-type magnetite ore in Yuanjiacun Iron Mine was processed adopting a processing technique consisting of high pressure grinding roll (HPGR) plus pre-concentration, followed by staged grinding and staged separation process. In order to reduce the grinding feed in the following concentration, HPGR was introduced for superfine crushing, and the obtained material in a size fraction of－3 mm was treated by a wet pre-concentration process, yielding 22% coarse tailings for discarding. And then a pre-concentration process resulted in the obtained concentrate with TFe grade up to around 36%. The work index for grinding before and after adopting HPGR was reduced from 11.91 kW·h/t to 11.11 kW·h/t. An experiment was performed with a flowsheet of HPGR plus pre-concentration followed by staged grinding and staged separation process, yielding 33.92% concentrate grading 67.13% TFe at 75.75% recovery, showing the ratio of concentration down to 2.27 from 2.95. It is concluded that the above-mentioned processing technique can provide an energy-conservation and consumption-reduction solution for large-scale and low-cost exploitation of similar low-grade magnetite mines.

A flowsheet consisting of pre-reduction magnetic roasting and magnetic separation was introduced to process a refractory weakly-magnetic iron ore from Jiangsu Province. The results show that after 60 min roasting at 750 ℃ by adding coal as a reducing agent at an amount of 3%, the pre-reduced iron ore was ground to a fineness of 75% -0.074 mm and then subjected to low intensity magnetic separation consisting of one roughing, one cleaning and one scavenging, resulting in a high quality iron concentrate grading 65.05% Fe at 96.19% recovery. The research results can provide technical reference for the mineral processing of refractory weakly-magnetic iron ore.

With sodium oxalate and D401 chelating resin as impurity removal materials, impurities of calcium and magnesium in crude lithium solution were removed by adopting a process consisting of chemical precipitation and ion exchange. In an experiment, reaction was firstly run at 25 ℃ for 60 min by adding sodium oxalate at an amount of 17.6 times the theoretical value, with pH of 6.5 and stirring rate of 300 r/min. And then, the lithium solution after filtration flowed through a resin column at a rate of 8 BV/h. As a result, the removal rates of calcium and magnesium reached 98.21% and 83.36%, respectively, and the lithium loss rate was 5.34%. The purified lithium solution was then recrystallized and purified, and the obtained lithium phosphate powder, with regular morphology and uniform dispersion, has its purity higher than 99.5%.