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.

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.

Based on an overview of progress in the research on iron recovery from high-iron red mud by alkaline electrolysis, the research findings at home and abroad are introduced in terms of principle and advantages of electrolysis, influencing factors for the process, and the obtained iron products. The advantages of red mud in the process for preparing metallic iron by electrolytic reduction of red mud in alkaline media at low temperatures are elaborated, and some problems to be solved in the future are proposed, which can provide references for innovation and development of this process.

Homogenization of backfill slurry cannot be automatically evaluated during stirring process. To solve this problem, the Yolo-v7 algorithm was improved by using the CBAM and the SPD-Conv, and the improved Yolo-v7 algorithm was then employed to automatically evaluate the homogenization of backfill slurry. The research results show that the improved Yolo-v7 algorithm demonstrates a significantly improved performance, with the accuracy, recall rate and mean average precision (mAP) all increased by 17.5, 28.8, and 32.4 percentage points respectively. Analysis of slurry parameter sensitivity indicates that slurry concentration is the principal factor affecting evaluation of slurry inhomogeneity, followed by the cement-tailings ratio. Moreover, the homogeneity of backfill slurry with high concentration can be significantly improved by prolonging stirring time.

After an introduction of the current status of gallium resources and analysis of the occurrence state of gallium in sodium aluminate solution, zinc smelting slag, tailings from vanadium extraction and coal-based solid waste, progress in the research of gallium extraction processes is elaborated and the adaptability, advantages and disadvantages of each technique are also summarized. Finally, based on the discussion of problems faced by the gallium industry and the development direction in the future, it is suggested that iteration and upgrading of gallium extraction processes should be vigorously promoted based on China's scale advantage in alumina industry.

An extraction system composed of P507 as the extractant and sulfonated kerosene as the diluent was used to separate zinc from a sulfuric acid leachate containing cobalt, manganese and zinc. The effects of key parameters on zinc separation efficiency were investigated, including the volume fraction of P507 in organic phase, saponification degree, pH value of feed, O/W ratio, reaction time, and reaction temperature. The results show that under the following optimal process conditions, including P507 at 20% by volume, saponification degree of 80%, feed pH of 5, O/W ratio of 1∶1, reaction time of 10 min and reaction temperature of 25 ℃, the extraction rate of Zn reaches 99.92%, while the extraction rates of Mn and Co are just 17.41% and 3.69%, respectively, presenting an excellent separation effect.

With nickel laterite ore from Morowali of Indonesia as the raw material, effects of various factors, including reaction time and temperature, stirring speed, acid-to-ore ratio, and slurry concentration, on the leaching rate were systematically investigated. Additionally, the leaching mechanism was explored by using the PHREEQC, a thermodynamic calculation software. The results show that after one hour reaction at 250 ℃, with acid-to-ore ratio of 300 kg/t and slurry concentration of 25%, the leaching rates of Ni and Co can reach 98.40% and 99.90%, respectively, while the leaching rates of Fe and Al are just 3.84% and 40.25%, respectively. The high selectivity of high pressure acid leaching (HPAL) is attributed by the factor that at high temperatures, not only the hydrolysis reaction of Fe and Al can be promoted, but also the pOH can be decreased from 14 to 10, which thereby inhibits leaching of Fe and Al. As the temperature rises to above 200 ℃, the number of H⁺ increases, which can ensure a sufficient amount of H⁺ for leaching of nickel and cobalt metals.

A process mineralogy study was conducted for a niobium-bearing ore in eastern Africa. Based on the study results, the raw ore was ground to－0.10 mm (100% passing size), and then pretreated with a process consisting of desliming, iron removal and zircon flotation for reducing the influence of slime, magnetite, and zircon on flotation performance. Under weakly acidic conditions, a closed-circuit niobium flotation test was carried out, with water glass and aluminum sulfate as a combined depressant, lead nitrate as an activator, and a self-developed chelating collector GYX₂, producing a niobium concentrate grading 26.87% Nb₂O₅ at 88.31% recovery, with a corresponding recovery of 80.86% from the raw ore. It is concluded that the niobium minerals in the ore can be effectively recovered.

The ore in the concentrator of Anhui Taiping Mining Company cannot reach the expected flotation effect. In order to address the problem, the properties and particle size composition of raw ore were analyzed, and condition tests were conducted to rule out the influence of reagent system and flotation conditions on the flotation performance. Based on a test on pre-desliming of the raw ore sample before grinding and the concentrate from pretreatment, it is found that the secondary slime has a significantly superior floatability than the primary slime, and too much primary slime is a contributing factor in the finally poor flotation performance. And based on a test on comparison of desliming, it is found that after pre-desliming process, the flotation effect of ore can be obviously improved, and the concentrate can have its final recovery improved by around 10 percentage points compared to that without pre-desliming process, due to significantly decreased slime content in concentrate and middlings. It is concluded that pre-desliming process is an effective approach for improving flotation performance of such kind of copper-bearing ore with serious sliming.

In order to solve the problems of fuzziness, randomness and uncertainty in optimization of mining method by traditional analytic hierarchy process (AHP), some measures, including introducing fuzzy theory, optimizing judgment scale and constructing optimal transfer matrix, were introduced to reduce subjective deviation and enhance the processing ability of uncertain information. Then, based on the combination of improved analytic hierarchy process (AHP) with the technique for order preference by similarity to ideal solution (TOPSIS), a comprehensive evaluation system with multi-level, multi-factor, multi-objective and multi-index characteristics was constructed and applied to the optimization of mining methods for a gently-inclined medium-thick phosphate ore body. The analysis shows that the relative closeness coefficients of those three mining schemes are 0.177 9, 0.822 5 and 0.648 3, respectively. Scheme 2, namely sublevel open stoping with subsequent backfill, is primarily determined as the suitable mining method for the mine. The numerical simulation of Scheme 2 also shows that the maximum surface subsidence in this scheme can be controlled within 72.94 mm, indicating that the mining operation at this stage brings little impact to the surface industrial zone on the mining site.