Porphyry copper deposits (PCDs) are some of the most important sources of copper (Cu) and molybdenum (Mo). Typically, the separation and recovery of chalcopyrite (CuFeS₂) and molybdenite (MoS₂), the major Cu and Mo minerals, respectively, in PCDs are achieved by two-step flotation involving (1) bulk flotation to separate Cu–Mo concentrates and tailings (e.g., pyrite, silicate, and aluminosilicate minerals) and (2) Cu–Mo flotation to separate chalcopyrite and molybdenite. In Cu–Mo flotation, chalcopyrite is depressed using Cu depressants, such as NaHS, Na₂S, Nokes reagent (P₂S₅ + NaOH), and NaCN, meaning that it is recovered as tailings, while molybdenite is floated and recovered as froth product. Although conventionally used depressants are effective in the separation of Cu and Mo, they have the potential to emit toxic and deadly gases such as H₂S and HCN when operating conditions are not properly controlled. To address these problems caused by the use of conventional depressants, many studies aimed to develop alternative methods of depressing either chalcopyrite or molybdenite. In this review, recent advances in chalcopyrite and molybdenite depressions for Cu–Mo flotation separation are reviewed, including alternative organic and inorganic depressants for Cu or Mo, as well as oxidation-treatment technologies, such as ozone (O₃), plasma, hydrogen peroxide (H₂O₂), and electrolysis, which create hydrophilic coatings on the mineral surface.

Flotation is a common mineral processing method used to upgrade copper sulfide ores; in this method, copper sulfide mineral particles are concentrated in froth, and associated gangue minerals are separated as tailings. However, a significant amount of copper is lost into tailings during the processing; therefore, tailings can be considered secondary resources or future deposits of copper. Particle–bubble collision efficiency and particle–bubble aggregate stability determines the recovery of target particles; this attachment efficiency plays a vital role in the selectivity process. The presence of fine particles in the flotation circuit is because of excessive grinding, which is to achieve a higher degree of liberation. Complex sulfide ores of markedly low grade further necessitate excessive grinding to achieve the maximum degree of liberation. In the flotation process, fine particles due to their small mass and momentum are unable to collide with rising bubbles, and their rate of flotation is very slow, further lowering the recovery of target minerals. This collision efficiency mainly depends on the particle–bubble size ratio and the concentration of particles present in the pulp. To overcome this problem and to maintain a favorable particle–bubble size ratio, different techniques have been employed by researchers to enhance particle–bubble collision efficiency either by increasing particle size or by decreasing bubble size. In this article, the mechanism of tailing loss is discussed in detail. In addition, flotation methods for fine particles recovery such as microbubble flotation, column flotation, nanobubble flotation, polymer flocculation, shear flocculation, oil agglomeration, and carrier flotation are reviewed, and their applications and limitations are discussed in detail.

Zinc plant leach residues (ZPLRs), particularly those produced using old technologies, have both economic importance as secondary raw materials and have environmental impacts because they contain hazardous heavy metals that pose risks to human health and the environment. Therefore, the extraction and recovery of these metals from ZPLRs has both economic and environmental benefits. In this study, we investigated the removal of lead (Pb) and zinc (Zn) from ZPLRs by alkaline (NaOH) leaching and the concurrent cementation of dissolved Pb and Zn using aluminum (Al) metal powder. The effects of the leaching time, NaOH concentration, solid-to-liquid ratio (S/L), and dosage of Al metal powder on the extraction of Pb and Zn were investigated. Pb and Zn removal efficiencies increased with increasing NaOH concentrations and decreasing S/Ls. The Pb and Zn removal efficiencies were 62.2% and 27.1%, respectively, when 2.5 g/50 mL (S/L) of ZPLRs were leached in a 3 M NaOH solution for 30 min. The extraction of Pb and Zn could be attributed to the partitioning of these metals in relatively more mobile phases—water-soluble, exchangeable, and carbonate phases—in ZPLRs. Around 100% of dissolved Pb and less than 2% of dissolved Zn were cemented in leaching pulp when Al metal powder was added. Minerals in the solid residues, particularly iron oxides minerals, were found to suppress the cementation of extracted Zn in leaching pulp, and when they were removed by filtration, Zn was recovered by Al metal powder via cementation.

The purpose of this study is to propose the flotation procedure of seafloor massive sulfide (SMS) ores to separate chalcopyrite and galena as froth and sphalerite, pyrite, and other gangue minerals as tailings, which is currently facing difficulties due to the presence of water-soluble compounds. The obtained SMS ore sample contains CuFeS₂, ZnS, FeS₂, SiO₂, and BaSO₄ in addition to PbS and PbSO₄ as Pb minerals. Soluble compounds releasing Pb, Zn²⁺, Pb²⁺, and Fe^2+/3+ are also contained. When anglesite co-exists, lead activation of sphalerite occurred, and thus sphalerite was recovered together with chalcopyrite as froth. To remove soluble compounds (e.g., anglesite) that have detrimental effects on the separation efficiency of chalcopyrite and sphalerite, surface cleaning pretreatment using ethylene diamine tetra acetic acid (EDTA) was applied before flotation. Although most of anglesite were removed and the recovery of chalcopyrite was improved from 19% to 81% at 20 g/t potassium amyl xanthate (KAX) after EDTA washing, the floatability of sphalerite was not suppressed. When zinc sulfate was used as a depressant for sphalerite after EDTA washing, the separation efficiency of chalcopyrite and sphalerite was improved due to deactivation of lead-activated sphalerite by zinc sulfate. The proposed flotation procedure of SMS ores—a combination of surface cleaning with EDTA to remove anglesite and the depression of lead-activated sphalerite by using zinc sulfate—could achieve the highest separation efficiency of chalcopyrite and sphalerite; that is, at 200 g/t KAX, the recoveries of chalcopyrite and sphalerite were 86% and 17%, respectively.

Porphyry Cu-Mo deposits, which are the most important sources of copper and molybdenum, are typically processed by flotation. In order to separate Cu and Mo minerals (mostly chalcopyrite and molybdenite), the strategy of depressing chalcopyrite while floating molybdenite has been widely adopted by using chalcopyrite depressants, such as NaHS, Na₂S, and Nokes reagent. However, these depressants are potentially toxic due to their possibility to emit H₂S gas. Thus, this study aims at developing a new concept for selectively depressing chalcopyrite via microencapsulation while using Fe²⁺ and PO₄³⁻ forming Fe^(III)PO₄ coating. The cyclic voltammetry results indicated that Fe²⁺ can be oxidized to Fe³⁺ on the chalcopyrite surface, but not on the molybdenite surface, which arises from their different electrical properties. As a result of microencapsulation treatment using 1 mmol/L Fe²⁺ and 1 mmol/L PO₄³⁻, chalcopyrite was much more coated with FePO₄ than molybdenite, which indicated that selective depression of chalcopyrite by the microencapsulation technique is highly achievable.