Solvent extraction is the most widely used method for separation and purification of rare earth elements, and organic extractants such as di(2-ethylhexyl) phosphoric acid (P204) and di(1-methyl-heptyl) methyl phosphonate (P350) are most commonly used for industrial applications. However, the presence of impurity ions in the feed liquid during extraction can easily emulsify the extractant and affect the quality of rare earth products. Aluminum ion is the most common impurity ion in the feed liquid, and it is an important cause of emulsification of the extractant. In this study, the influence of aluminum ion was investigated on the extraction of light rare earth elements by the P204-P350 system in hydrochloric acid medium. The results show that Al³⁺ competes with light rare earths in the extraction process, reducing the overall extraction rate. In addition, the Al³⁺ stripping rate is low and there is continuous accumulation of Al³⁺ in the organic phase during the stripping process, affecting the extraction efficiency and even causing emulsification. The slope method and infrared detection were utilized to explore the formation of an extraction compound of Al³⁺ and the extractant P204-P350 that entered the organic phase as AlCl[(HA)₂]₂P350_(o).

Zinc-leaching residue (ZLR) is a strongly acidic hazardous waste; it has poor stability, high heavy metal levels, and releases toxic elements into the environment. ZLR has potential as a valuable resource, because it contains elevated levels of zinc and silver. In this paper, the recovery of zinc (Zn) and silver (Ag) from ZLR wastes from zinc hydrometallurgy workshops using water leaching followed by flotation was studied. During water leaching experiments, the zinc and copper recovery rates were 38% and 61%, respectively. Thereafter, various flotation testing parameters were optimized and included grinding time, reagent dosages, pulp density, flotation time, and type of adjuster. Experimental results demonstrated this flotation method successfully recycled Ag and Zn. A froth product containing more than 9256.41 g/t Ag and 12.26% Zn was produced from the ZLR with approximately 80.32% Ag and 42.88% Zn recoveries. The toxicity characteristic leaching procedure (TCLP) results indicated the water-leaching flotation process not only recycled valuable metals such as zinc and silver in zinc-containing hazardous wastes but lowered the hazardous waste levels to those of general wastes and recycled wastes in an efficient, economical, and environmentally friendly way.

Copper slag (CS), a prominent industrial by-product derived from copper sulphide concentrate smelting, contains valuable copper minerals. Mineralogical analysis revealed 0.56% arsenical copper/chalcopyrite, 1.29% chalcocite, 0.25% cuprite, and 0.24% bornite as the primary copper-bearing phases. Given the rapid depletion of readily exploitable copper resources, CS has emerged as a significant source for copper production. A systematic investigation was conducted to establish optimal operational parameters for classification, grinding, and flotation. This study encompassed diverse scenarios and laboratory tests, with a particular focus on reagent application, including types and consumption rates. The findings demonstrated that Na₂S is used to activate copper oxide minerals that are difficult to recover after a period of grinding, increasing the recovery rate of CS from 53.92% to 87.58%, and the two-stage grinding process is beneficial for precise control of the particle size, considering both energy efficiency and mineral recovery rate. Upon conducting a comprehensive investigation into the closed-circuit flotation process for copper recovery from CS under optimal conditions, it was determined that the copper concentrate achieved a grade of 21.24% and a recovery rate of 36.31%. Additionally, the grade of copper in the flotation tailings was diminished to 0.20%. The synchronous enrichment of copper significantly curtails total emissions in pyrometallurgical processing, mitigating pollution from CS flotation tailings. This study offers a scientific foundation for sustainable CS tailing management, promoting resource recovery and reduced environmental impact.

Tennantite (Cu₁₂As₄S₁₃) and enargite (Cu₃AsS₄) are two important minerals that simultaneously contain copper and arsenic. Detailed studies of their structure and properties are crucial for understanding their oxidation, flotation, and leaching. This study investigates the crystal structures, electronic properties, and reactivity of these two copper-arsenic minerals from the perspectives of atomic bonding, charge, density of states, and d-orbital splitting. The results indicate that tennantite is a crystal with mixed Cu valence states of +2 and +1 (predominantly +1), while the Cu in enargite is in the +1 state. The valence state of As in tennantite (+3) is lower than that in enargite (+5). Orbital energy level calculations show that the energy gaps between the copper d-orbitals are small in both minerals, indicating strong electron delocalization and, consequently, strong covalent character in the crystals, which is also confirmed by Mulliken bond population calculations. The presence of arsenic is the reason for the enhanced covalency. It is noteworthy that tennantite exhibits stronger covalency. The Cu 3d and As 4p electrons in tennantite are more electronically active than those in enargite. In tennantite, the strong d-electron delocalization caused by d-p hybridization between Cu and S leads to similar 3d electronic properties between 3-coordinated and 4-coordinated Cu. The energies of the five d-orbitals of the 4-coordinated Cu in enargite are lower than those of the 4-coordinated Cu in tennantite, which may affect the ability of Cu 3d electrons to enter the empty orbitals of S atoms in sulfur-containing collectors to form π back-bonding, thereby reducing the collecting ability of enargite. On the other hand, the splitting energy of the 4-coordinated Cu 3d orbitals in enargite is significantly smaller than that in tennantite, making the structure less stable and, thus, potentially more prone to dissolution.

Direct cyanide leaching of high-carbon gold ores typically results in poor gold extraction due to the “preg-robbing” effect of carbonaceous matter. This study investigates a refractory high-carbon gold ore from Laos. Based on detailed process mineralogy that identified the occurrence state of gold and the distribution of harmful impurities, a combined process consisting of flotation pre-concentration, oxidative roasting, and a Mill–Leach Coordination (MCL) process was developed. The MCL process, which integrates simultaneous grinding and leaching, continuously renews the particle surface and eliminates product-layer diffusion resistance, thereby significantly enhancing gold dissolution efficiency. Kinetic studies demonstrate that the leaching reaction follows the Avrami model and is controlled by internal diffusion, with a derived semi-empirical kinetic equation. This research provides a novel and efficient technical route for treating high-carbon gold ores and offers theoretical guidance for industrial scale-up.

Phosphogypsum (PG) and phosphorus tailings (PT) are bulk solid wastes generated by the phosphorus chemical industry whose stockpiling poses significant environmental risks and represents a waste of resources. To achieve the goals of “treating waste with waste” and large-scale disposal, this study proposes a technical pathway involving the co-pyrolysis of phosphogypsum and phosphorus tailings to produce artificial soil-like materials. The effects of raw material ratio, pyrolysis temperature and duration, and biomass addition on the speciation transformation, leaching toxicity, and matrix characteristics of phosphorus (P) and fluorine (F) in the products were systematically investigated. Characterization techniques, including X-ray diffraction (XRD), Fourier transform infrared spectroscopy (FTIR), and scanning electron microscopy (SEM), were employed to elucidate the synergistic immobilization mechanism. The results indicate that under optimized conditions (PG:PT mass ratio of 6:4, pyrolysis temperature of 800 °C, duration of 2–3 h, and biomass addition of 20%–30%), the active forms of harmful elements in the product were significantly reduced. The proportion of water-soluble fluorine decreased from ~39% in raw phosphogypsum to less than 3%, with apatite phosphorus becoming the dominant form of phosphorus. Mechanistic studies reveal that the immobilization process follows a “multi-pathway synergy” mechanism: thermal activation promotes the in situ formation of thermodynamically stable fluorapatite through the reaction of Ca²⁺, PO₄³⁻, and F⁻ (chemical fixation); iron/aluminum oxides in phosphorus tailings and the biochar derived from added biomass provide adsorption sites for surface complexation (physicochemical fixation); and the melting of silicon–aluminum components forms an amorphous silicate network that physically encapsulates pollutant microcrystals. This study provides crucial theoretical foundations and process parameters for the synergistic disposal and soil-like resource utilization of phosphogypsum and phosphorus tailings, demonstrating significant environmental and economic benefits.