This research investigates the hydrometallurgical processing of molybdenum middlings extracted from copper-molybdenum ore in the Shatyrkul-Zhaysan cluster in Kazakhstan. A molybdenum intermediate obtained after selective flotation of the copper-molybdenum concentrate was used, with a recovery of 0.07%, a molybdenum content of 22.23% and an extraction of 74.91%. The mineralogical analysis shows molybdenite and chalcopyrite as the main minerals. In experiments, atmospheric leaching with nitric acid in single-stage and two-stage countercurrent processes was investigated to optimize molybdenum extraction and reduce acid consumption. The optimum conditions obtained were: 300 g/L nitric acid, 100 g/L sulfuric acid, 90?C temperature and 2 hours leaching time in the single-stage leaching, which extracted 98.8% molybdenum. The two-stage leaching under optimized conditions allows the extraction of 94.3% molybdenum in solutions with lower residual acidity (0.89 g-eq/L) and redox potential (550 mV) without reducing the extraction of valuable component. Molybdenum extraction reaches 94.3% in the subsequent solvent extraction stage from two-stage leaching solutions. The final product, calcium molybdate with a molybdenum content of 46.83%, meets commercial grade specifications. This research demonstrates an effective process for hydrometallurgical production of commercial grade calcium molybdenite from copper-molybdenum ore, with high molybdenum recovery, reduced acid consumption through two-stage leaching, and minimal hazardous discharges.

This paper presents an analysis of the current state of processing lead–zinc ores from the Koskudyk deposit (Kazakhstan). At present, polymetallic ores are being extracted from the Ridder-Sokolnoye, Zyryanovskoye, Maleevskoye, and Achisai deposits. However, the reserves of rich and easily beneficiable ores are being depleted, and the supply of raw materials from the developed deposits does not exceed 25 years. As a result, more complex and difficult-to-enrich oxidized and mixed ores are being involved in production, and the extraction of non-ferrous metals from these ores presents a significant technological challenge. The most effective method for enriching oxidized polymetallic ores is flotation with preliminary sulfidization. Laboratory studies were conducted on a sample of oxidized lead–zinc ore from the Koskudyk deposit, which contains 79.69% oxidized lead compounds and 84.72% oxidized zinc compounds. This study examines the effect of sulfidization using sodium sulfide and determines the oxidative–reductive potential (ORP) levels for various reagent dosages. The experiments demonstrated that a sodium sulfide dosage of 700 g/t at an ORP of −200 mV yields the most effective lead flotation, resulting in a lead recovery of 50.07%. Zinc recovery remained relatively unchanged across all tests, confirming the limited response of oxidized zinc minerals under the applied sulfidization conditions. The highest beneficiation efficiency was achieved within the ORP range of −160 to −200 mV, beyond which lead recovery began to decline. The findings underscore the importance of optimizing ORP to ensure the formation of a stable sulfide film on mineral surfaces and efficient collector attachment. These results provide practical guidance for improving flotation performance of oxidized ores and demonstrate the need for additional activation strategies in zinc recovery.

This study examines the impact of regrinding on the interfacial properties of sulfide minerals and the flotation performance of weathered copper–porphyry tailings. The feed material is characterized by a low copper grade (0.17%) and a high proportion of oxidized species (53.84%), which contribute to its inherent chemical stability and poor flotation kinetics. The findings indicate that regrinding serves a dual role: facilitating the liberation of mineral intergrowths and inducing mechanical surface renewal. This renewal is characterized by a significant decrease in the oxidation–reduction potential (ORP) and an intensification of the surface reactivity. Experimental results identify an optimal grinding fineness of 77%–81% passing −0.045 mm, yielding a copper recovery of 16.26% in the absence of a sulfidizing agent. The integration of sodium sulfide (400 g/t) with regrinding significantly enhances recovery to 36.37%, driven by the establishment of a reducing environment (ORP ≈ −150 mV) and the chemisorption-mediated activation of mineral surfaces. While ultrafine grinding (90%–100% passing −0.045 mm) further increases recovery to 51.47%, it is accompanied by deleterious sliming effects and a subsequent loss of process selectivity. The study confirms that mechanical surface rejuvenation and the optimization of electrochemical conditions are critical for improving the processing efficiency of anthropogenic resources, providing a theoretical framework for establishing rational beneficiation regimes.

This paper examines the mineral and chemical composition of converter slags generated by Balkhash Copper Smelter, as well as the forms in which copper ispresent in them. The paper describes the results of flotation of the above product, which was prior subjected to fine crushing in jaw and hammer crushers. The paper features a comparative analysis that looks at the size distribution and the distribution of the key valuable components by grain size of the slag pre-crushed in jaw and hammer crushers. It is shown that the different breaking techniques realized by these crushers influence the size distribution of the crushed product, the share of particles smaller than 0.071 mm and the distribution of copper by fraction. The authors looked at the grinding kinetics of the copper smelter slags pre-crushed in the above specified crushers and compared the share of –0.071 mm particles, as well as that of –0.045 mm particles (which are the finest fraction). It was established that the optimum hammer crushing duration is 45 min. A longer duration leads to overgrinding. At the same time for the jaw crusher, the overgrinding effect and the accumulation of the –0.045 mm particles become expressed if the crushing cycle lasts for 60 min. Flotation of the pre-crushed converter slags resulted in a quality concentrate containing 27.5 to 39.0 wt % copper (depending on the slag crushing duration). In all cases, the recovery of copper was every time higher (by 5% on average) for slags pre-crushed in the hammer crusher. It is pointed out that impact crushing of copper smelter slags significantly intensifies the flotation process. This should be taken into consideration when optimizing the processing schemes.This research was funded by the Russian Science Foundation; Agreement No. 22-27-00526.

This paper presents an analysis of the current state of processing lead-zinc ores from the Koskudyk deposit (Kazakhstan). At present, polymetallic ores are being extracted from the Ridder-Sokolnoye, Zyryanovskoye, Maleevskoye, and Achisai deposits. However, the reserves of rich and easily beneficiable ores are being depleted, and the supply of raw materials from the developed deposits does not exceed 25 years. As a result, more complex and difficult-to-enrich oxidized and mixed ores are being involved in production, and the extraction of non-ferrous metals from these ores presents a significant technological challenge. The most effective method for enriching oxidized polymetallic ores is flotation with preliminary sulfidization. Laboratory studies were conducted on a sample of oxidized lead-zinc ore from the Koskudyk deposit, which contains 79.69% oxidized lead compounds and 84.72% oxidized zinc compounds. This study examines the effect of sulfidization using sodium sulfide and determines the oxidative-reductive potential (ORP) levels for various reagent dosages. The results showed that increasing the sodium sulfide dosage to 700 g/t and achieving an ORP of −200 mV led to a 50.07% lead extraction, while the quality and quantity of zinc extraction remained stable. Determining the optimal ORP level significantly improves the efficiency of the enrichment process, enhancing both qualitative and quantitative indicators and simplifying the regulation of the technological process.

The flotation of oxidized lead–zinc ores is challenging due to the low floatability of oxidized minerals and their weak interaction with conventional reagents. This study investigates the influence of pulp electrochemical parameters, namely redox potential (Eh) and pH, on the flotation kinetics of oxidized lead–zinc ore from the Koskuduk deposit. The results showed that sodium sulfide selectively activates lead-bearing minerals, increasing Pb recovery to 40.74%, while Zn recovery remains low at approximately 12%. In contrast, the polysulfide–lime system S:CaO:H₂O provides more uniform and stable sulfidization of oxidized mineral surfaces, increasing recovery to 65.10% for Pb and 56.89% for Zn. The highest recoveries were achieved within an Eh range of −120 to −180 mV at pH 11–12. Kinetic studies demonstrated that the main contribution to metal recovery occurs during the first 2–6 min of flotation. These results indicate that flotation efficiency is controlled not only by reagent type but also by the electrochemical state of the pulp, confirming that calcium polysulfide is a promising alternative sulfidizing reagent for processing oxidized lead–zinc ores.