An experimental research was conducted on efficient recovery of ilmenite from Panzhihua City by magnetic fluid coupling high-gradient magnetic separation. With MnCl₂ concentration of 40%, magnetic induction intensity of 0.6 T, and pulsation frequency of 400 r/min, the ilmenite was processed by magnetic fluid coupling high-gradient magnetic separation, including one roughing and three cleaning, resulting in a titanium concentrate grading 46.89% TiO₂ at 38.30% recovery. It is shown that the obtained indices are comparable to those obtained from industrial production by adopting magnetic separation combined with flotation process.

The ultra-fine low-grade calcite-containing apatite concentrate that comes from Hebei Province cannot meet the grade requirement and is difficult to be utilized. For solving this problem, an experimental research was carried out to improve the concentrate grade. A depressant of CD-8 was adopted for calcite, the dominate gangue mineral affecting the grade of apatite concentrate, and a collector of CK-P with high selectivity was adopted for apatite. A closed-circuit experiment was performed with a direct flotation flowsheet, consisting of one roughing, one scavenging and three cleaning, with middlings sequentially returned. As a result, a phosphate concentrate with P₂O₅ grade of 32.02% was finally obtained with a recovery of 70.92%. This flowsheet was applied to industrial production with normal operation and stable flotation indicators, indicating that efficient utilization of such ultra-fine and refractory phosphorus resource was actualized.

Based on experimental studies on separation and quality improvement for a boron-bearing iron ore, a scheme for boron extraction and sulfur reduction was proposed, and a metallurgical process combined with mineral processing was determined, which consists of activation roasting, grinding, alkali leaching and magnetic separation. After activation roasting at 650 ℃ for 90 min in a neutral atmosphere, the obtained roasted ore was ground to a fineness of -0.075 mm 83.28%, and then leached for 60 min at 100 ℃, with liquid-to-solid ratio of 3∶1, leaching reagent of NaOH at a concentration of 15%, and stirring speed of 180 r/min. A boron-bearing mother liquor was obtained, with leaching rate of B₂O₃ at 94.62%. The leaching residue was subjected to conventional magnetic separation, resulting in an iron concentrate grading 54.94% TFe, containing 0.24% B₂O₃ and 0.077% S. It is concluded that boron can be fully recovered and the quality of iron concentrate can be also improved by adopting this processing scheme.

As for the leaching solution of gold with copper(II)-ethylenediamine-thiosulfate, an adsorbent (TTAB@Sch) was prepared by modifying schwertmannite (Sch) with tetradecyltrimethylammonium bromide (TTAB), and applied to adsorb arsenic therein. The effects of Cu²⁺ concentration, ethylenediamine concentration, S₂O₃^2- concentration, solution pH and initial arsenic mass concentration on the adsorption of arsenic by TTAB@Sch were explored. The results show that with the addition of FeSO₄·7H₂O, KMnO₄ and TTAB at amounts of 10, 1 and 2 g/L respectively, the prepared TTAB@Sch exhibits optimal adsorption behavior. As for the leaching solution with Cu²⁺ concentration of 5 mmol/L, ethylenediamine concentration of 12.5 mmol/L, initial arsenic mass concentration of 12.29 mg/L, and pH of 10, the adsorption rate of arsenic therein by TTAB@Sch can reach 87.2%. Furthermore, FT-IR and XPS analyses indicated that arsenic was likely adsorbed on TTAB@Sch by forming Fe—O—As.

Based on an introduction of nickel resource and characteristics of lateritic nickel ore, the technical features and application of traditional hydrometallurgical processes (such as high-pressure acid leaching (HPAL), reduction roasting-ammonia leaching) and pyrometallurgical processes (such as rotary kiln electric furnace (RKEF), blast furnace smelting process for ferronickel production, and rotary kiln direct reduction-magnetic separation) are summarized. A new process of suspension roasting pre-reduction electric furnace (SRPEF) is specially introduced. Finally, the prospects for efficient exploitation and clean extraction technologies for lateritic nickel ore are discussed.

To address the issues of low purity and high consumption of alkali and acid during chemical purification of graphite, Luobei spherical graphite was taken to explore the occurrence state of impurities therein and to perform tests on its purification by alkali-acid method. A novel alkali-acid method was developed, which consists of stirring with alkali under negative-positive pressure and pressure acid leaching. This purification process is conducted in the following steps with the optimized conditions: firstly leave graphite in a vacuum for 35 min with vacuum degree of -0.15 MPa, and then stir for 40 min under pressure of 0.19 MPa at 120 ℃ by adding NaOH at a concentration of 40%, with NaOH solution and spherical graphite in a liquid-to-solid ratio of 1.0 mL/g；after 120 min roasting at 550 ℃, acid leaching with HCl at a concentration of 2.5 mol/L is performed under pressure of 0.16 MPa at 118 ℃ for 90 min, with HCl solution and graphite in a liquid-to-solid ratio of 1.0 mL/g. As a result, the fixed carbon content in the purified graphite is up from 95.35% to 99.96%.

In order to recycle the waste circuit boards, the acid leachate with a copper content of 34.33 g/L was treated by adopting solvent extraction to recover the copper therein. The effects of extractant type, extraction time, phase ratio, extractant concentration, and extraction temperature on copper extraction efficiency were all investigated. The results show that with AD100 as the copper extractant, phase ratio (V_O/V_A) at 1∶3, and extractant concentration of 60%, an oscillating extraction at room temperature of 25 ℃ for 10 min can result in the copper extraction rate of 96.62%. Then, a back-extraction process is adopted to treat the organic phase with oxalic acid, producing copper oxalate, which can be calcined to obtained copper oxide products.

With spent lithium manganese iron phosphate cathode powder as raw material, a leaching system of DL-malic acid combined with sodium persulfate was employed to selectively recover Li, Mn and Fe elements. The effects of factors, including DL-malic acid concentration, molar ratio of Na₂S₂O₈ to LiFe_0.4Mn_0.6PO₄, liquid-to-solid ratio, leaching temperature, and leaching time on the leaching rates of Li, Fe, Mn and P were investigated. Then, the left leaching residue was analyzed by using SEM and XRD, and the leachate was analyzed by FTIR. Results show that with DL-malic acid at a concentration of 0.4 mol/L, Na₂S₂O₈ and LiFe_0.4Mn_0.6PO₄ at a molar ratio of 0.4, liquid-to-solid ratio of 15 mL/g, 3 h leaching process at 30 ℃ can bring the leaching rates of Li, Fe, Mn and P at 98.76%, 2.37%, 96.89% and 52.45%, respectively, and also Li, Fe, Mn and P in the leachate in a mass concentration of 2.98, 0.21, 12.66 and 6.52 g/L, respectively. It is found that iron phosphate is the dominant component in the leaching residue. During the leaching process, DL-malic acid and Mn²⁺ form a complex that can be solved in leachate, leading to efficient separation between iron and manganese.

Effects of roasting temperature and heating rate on the crystalline form evolution of calcium arsenate during heat treatment were investigated. The heat-treated calcium arsenate was leached by strong acids (sulfuric acid, nitric acid, hydrochloric acid, aqua regia), and the relationship between the crystalline form of calcium arsenate and arsenic concentration in leaching solution was investigated. The results show that calcium arsenate undergoes significant crystalline form transition during roasting process, and higher roasting temperature can result in higher degree of crystallization. The heat-treated calcium arsenate is leached with acids and it is shown that less arsenic is leached out as the roasting temperature rises. After calcium arsenate is subjected to roasting at 500 ℃, the concentration of arsenic in the leaching solution can meet GB 5085.3—2007. After calcium arsenate is subjected to a roasting process at 900 ℃ and above, the concentration of arsenic in the leaching solution satisfied GB 3838—2002.

A stockpile of hazardous waste aluminum ash generated during aluminum smelting has posed serious environmental pollution risks. In order to solve this, a novel approach is proposed, in which the nitrogen-enriched aluminum ash is used in the steelmaking process to promote microalloying with vanadium and nitrogen. Aluminum ash is heated and stir-fried in a nitrogen atmosphere, resulting in its reaction with nitrogen gas, thereby increasing the nitrogen level therein. The effects of reaction temperature, aluminum content and particle size on the nitrogen enrichment of aluminum ash were investigated, and the nitrogen-enriched aluminum ash was characterized by XRD and SEM-EDS. The results show that under the suitable conditions for nitrogen enrichment, including reaction temperature of 880 ℃, initial aluminum content of 45% (aluminum powder), and nitrogen flow rate at 4 L/min, the nitrogen content in the aluminum ash can be improved from 1.18% to 11.90%. During this process, metallic aluminum in the ash reacts with nitrogen gas to form AlN on the surface of alumina particles.