To comprehensively utilize molybdenum tailings, experimental researches were systematically carried out for the molybdenum tailings from a concentrator of Jinduicheng Molybdenum Mine. Based on process mineralogy study, valuable components in the tailings were determined to be stepped recovered by flotation. Firstly, by adopting kerosene combined with No.2 oil, a flotation process consisting of one roughing and two scavenging resulted in 21.23% Mo recovery to the concentrate grading 0.47% Mo, and tailings with Mo grade of just 0.008 7%. Secondly, a process of pre-desliming followed by flotation with TB-3 collector was applied to remove mica, resulting in the flotation tailings with significantly reduced mica content, which laid a foundation for subsequent separation. Finally, such flotation tailings were processed by adopting fluorine-free alkaline flotation for separating feldspar and quartz therein without regrinding. As a result, feldspar products with a yield of 29.23% and K₂O and Na₂O contents of 7.34% and 1.01%, respectively, and quartz products with a yield of 27.87% and SiO₂ grade of 83.14% were successfully produced. It is shown that such molybdenum tailings exhibit great potential for comprehensive recovery. It is concluded that the proposed combined process flow consisting of re-flotation of molybdenum, mica removal, and separation of feldspar and quartz by fluorine-free alkaline flotation is technically feasible for simultaneous recovery of metallic and non-metallic resources.

Conventional hydroxamic acid collectors suffer from weak collecting capacity and high dosage consumption. To address these drawbacks, 4-(chloromethyl) benzohydroxamic acid was synthesized, and its flotation performance for wolframite was systematically investigated. With methyl 4-(chloromethyl) benzoate, hydroxylamine hydrochloride and sodium hydroxide as raw materials, the optimal synthesis conditions were determined as follows:methyl 4-(chloromethyl) benzoate, hydroxylamine hydrochloride and sodium hydroxide in a molar ratio of 1∶1.2∶2.35, reaction temperature of 35 ℃, reaction time of 4 h, and pH of 5 at the terminal acidification. Tests on flotation of wolframite indicate that 4-(chloromethyl) benzohydroxamic acid outperforms benzohydroxamic acid in collecting ability. To achieve equivalent flotation indices, the required dosage of the synthesized reagent is lower than that of benzohydroxamic acid.

To improve the electrochemical performance of aluminum-air batteries, Br^- and I^- with weak electronegativity were adopted to regulate the hydrogen-bond structure of electrolyte and modify the electrode-electrolyte interface, so as to suppress hydrogen evolution corrosion of aluminum anode. The results indicate that the introduction of Br^- and I^- into electrolyte can significantly reduce the hydrogen evolution rate and self-corrosion rate of aluminum anodes, and remarkably improve the electrochemical performance of both aluminum anodes and aluminum-air batteries. After adding 3 mol/L KBr and 5 mol/L KI into KOH electrolyte, the hydrogen evolution rate and self-corrosion rate of aluminum anodes decrease to 13.79 mL/(cm²·h) and 13.5 mg/(cm²·h), 9.54 mL/(cm²·h) and 11.5 mg/(cm²·h), respectively. The corresponding corrosion inhibition efficiencies are 42.4% and 50.9%. Furthermore, the discharge specific capacities of the aluminum-air batteries can reach 1 167.1 mAh/g and 1 674.4 mAh/g at 25 mA/cm².

Aiming at the differences in physic-chemical properties of various iron concentrates and the difficulty in fully eliminating raw-material fluctuation by conventional dry blending, a technique of wet mixing combined with intensive mixing was adopted, and its effect on mixture homogeneity and pellet properties were investigated. The results show that wet mixing for 3 min can greatly improve the homogeneity of TFe, FeO and-0.074 mm fraction content in the mixed iron concentrates. After 3 min wet mixing and 40 s intensive mixing, the mixture has each homogeneity index basically equivalent to that after 60 s independent intensive mixing. The obtained green pellets have the compressive strength and drop strength improved by 12.17% and 13.95%, respectively, and preheated pellets and fired pellets have the compressive strength up by 7.64% and 10.63%, respectively, compared to that after raw material is just processed with 40 s intensified mixing without wet mixing. Meanwhile, the fluctuations in various pellet properties are remarkably reduced. It is concluded that this process can improve the homogeneity of various iron concentrates at the source and shorten the intensive mixing time, which provides technical support for quality improvement and stable operation in pelletization.

Aiming at the bottleneck that NH₃ and SO₂ generated during ammonium sulfate pyrolysis tend to spontaneously recombine and are difficult to be separated efficiently, an integrated process consisting of crystallization, pyrolysis and recycling, with CuO as the oxygen carrier in chemical looping was proposed. Experimental results indicate that the ammonia release rate exceeds 96% at the ammonia releasing stage after reaction at 400 ℃, with ammonium sulfate and copper oxide in a molar ratio of 1∶1.5. While at the sulfur releasing stage, after 90 min reaction at 750 ℃, sulfur release rate is over 98%. It is found that after five cycles of CuO-based chemical looping pyrolysis, both the ammonia-release and sulfur-release rates remain higher than 95%, and CuO also can be regenerated via air calcination. This process provides a feasible technical route for resource utilization of nitrogen-containing wastewater.

An optimized two-step separation-titration method was adopted to determine calcium fluoride content in a complex low-grade fluorite ore. Firstly, the sample was leached with 2% dilute hydrochloric acid for 150 min under magnetic stirring to efficiently separate interfering minerals such as calcite, dolomite and apatite, showing the leaching rates exceeding 98%. The residue was further leached with 80 g/L aluminum trichloride solution, and the calcium fluoride content was subsequently quantified via EDTA titration. Given that approximately 10% of fluorite would be leached by 2% hydrochloric acid, a correction coefficient (K=1.10) was introduced to calibrate the test results. The experimental results demonstrate that the relative errors for samples with various calcium fluoride contents determined by this method are within the permissible range (±0.25%), showing excellent consistency with theoretical blending values. The determination results are basically consistent with data obtained by Mineral Liberation Analyzer (MLA) and total fluorine content testing. It is concluded that this method effectively addresses the difficulty in accurate determination of calcium fluoride in complex low-grade fluorite ore, and can provide a reliable reference for mineral resource evaluation and optimization of beneficiation and metallurgical processes.

To optimize the blending of copper slag and copper concentrate in copper smelting process, thermodynamic analysis, phase-diagram calculation and experimental research were combined to systematically investigate the effects of blending ratio on phase evolution, sintering characteristics and sulfur separation behavior. The results show that an increase in the copper slag ratio from 40% to 60% leads to a remarkable rise in the fraction of oxide phases and a drop in the fraction of sulfide phases in the system, as well as higher liquid-phase-forming capacity. The structure of the sintered product gradually changes from loose and porous to dense and continuous, and reaches its most stable state when the copper slag ratio ranges from 50% to 60%. A higher copper slag ratio facilitates the formation of Fe-Ca-Si composite oxides, which greatly enhances the structural stability of sintered products. With a rise in oxygen concentration from 40% to 60%, sulfur volatilization capacity is substantially improved, and the sulfur volatilization efficiency rises by 11 percentage points.

Ammonium sulfate roasting - water leaching process was adopted to synergistically leach lithium and aluminum from coal ash, and the effects of mass ratio of ammonium sulfate to fly ash, roasting temperature, roasting time, liquid-to-solid ratio, leaching temperature and leaching time on leaching rates of lithium and aluminum were also investigated. The results show that under the optimized conditions:ammonium sulfate and fly ash in a mass ratio of 4∶1, roasting temperature of 400 ℃, roasting time of 2 h, liquid-to-solid ratio of 4∶1, leaching temperature of 80 ℃ and leaching time of 30 min, the leaching rates of lithium and aluminum reach 86.77% and 78.04%, respectively. Mechanism analysis indicates that mullite, the host phase of lithium, reacts with ammonium bisulfate (the thermal decomposition product of ammonium sulfate) at 400 ℃ to generate soluble ammonium aluminum sulfate and aluminum sulfate, resulting in destruction of mullite crystal structure. After being released, lithium reacts with ammonium bisulfate to form lithium sulfate, which enters the aqueous phase during water leaching, thus realizing efficient synergistic leaching of lithium and aluminum.

An improved Mask R-CNN based method for precursor particle detection is proposed. Multi-branch feature extraction is employed to capture particle features for feature fusion, and detection results of precursor particles are output through a dual-branch output module. The experimental results demonstrate that the proposed method achieves a detection accuracy of 93.2% for precursor particles, with an inference time of 15.1 ms per image. The model is characterized by both high accuracy and real-time inference efficiency, which can provide critical technical support for the production of highly-consistent precursor particles.

Modified metal-organic framework (MOF), UIO-66-TFA, was synthesized via hydrothermal reaction of zirconium tetrachloride, with terephthalic acid as the ligand and trifluoroacetic acid as the modulator. The characterization results reveal that the as-prepared modified MOF exhibits highly dispersed and uniform nanostructure. With adsorption time of 50 min, initial solution pH of 8, and reaction temperature of 40 ℃, an addition of 0.5 mg/mL UIO-66-TFA can result in the adsorption capacities toward Pb²⁺, Zn²⁺ and Mn²⁺ reaching 693.5 mg/g, 419.9 mg/g and 367.1 mg/g, respectively in the aqueous solution. After five consecutive adsorption-desorption cycles, UIO-66-TFA can still retain more than 88% of its initial adsorption capacity for Pb²⁺, Zn²⁺ and Mn²⁺.