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) contain significant amounts of metal compounds of lead (Pb), zinc (Zn), iron (Fe), etc., hence, they are considered as a secondary source of metals. On the other hand, ZPLRs are regarded as hazardous materials because they contain heavy metals that pollute the environment. Resources and environmental concerns of ZPLRs were addressed in this study by removing/recovering Pb and Zn using a concurrent dissolution and cementation technique. To cement the dissolved Pb and Zn in leaching pulp, zero-valent aluminum (ZVAl) was added during ZPLRs leaching in the hydrochloric (HCl)–sodium chloride (NaCl) solution. The resulting cemented metals were agglomerated and separated by sieving. Lead removal increased with increasing both NaCl and HCl concentrations. However, when ZVAl was added, significant Pb removal was achieved at a low concentration. Zinc was not cemented out of the pulp using ZVAl and its recovery from ZPLRs was dependent on the HCl concentration only. By applying a concurrent dissolution and cementation technique, both Pb and Zn were removed using a low concentration of NaCl, and most importantly Pb—the most toxic metal in ZPLRs—was captured and separated before the solid-liquid separation, hence, eliminating the need for extensive washing of the generated residues to remove the inherent residual solution.

Rare earth minerals (REMs) contain rare earth elements (REEs) that are important in modern technologies due to their unique magnetic, phosphorescent, and catalytic properties. However, REMs are not only non-renewable resources but also non-uniformly distributed on the Earth’s crust, so the processing of REE-bearing secondary resources via recycling is one potential route to ensure the long-term sustainability of REE supply. Flotation—a method that separates materials based on differences in their surface wettability—is a process applied for both mineral processing and recycling of REEs, especially when the particles are fine and/or a high-purity product is required. In this review, studies about rare earth flotation from 2012 to 2021 were systematically reviewed using the PRISMA guideline. It was found that most REM flotation research works focused on finding better collectors and depressants while, for recycling, studies on advanced flotation techniques like froth flotation, ion flotation, solvent sublation, electroflotation, and adsorbing colloid flotation with an emphasis on the recovery of dissolved REEs from aqueous solutions dominated.

The global demand for rare earth elements (REEs) is expected to increase significantly because of their importance in renewable energy and clean storage technologies, which are critical for drastic carbon dioxide emission reduction to achieve a carbon-neutral society. REE ore deposits around the world are scarce and those that have been identified but remain unexploited need to be developed to supply future demands. In this study, the Khalzan Buregtei deposit located in western Mongolia was studied with the aim of upgrading low-grade REE ore via magnetic separation techniques. The total REE content in this ore was ~6720 ppm (~3540 ppm light REE (LREE) + ~3180 ppm heavy REE (HREE)) with bastnaesite, pyrochlore, synchysite, and columbite-(Fe) identified as the main REE-bearing minerals. As the particle size fraction decreased from −4.0 + 2.0 mm to −0.5 + 0.1 mm, the recovery by dry high-intensity magnetic separation (DHIMS) increased from 20% to 70% of total rare earth oxide (TREO) while the enrichment ratio reached 2.8 from 1.3. Although effective, gangue minerals such as quartz and aluminosilicates were recovered (~22%) due most likely to insufficient liberation. Meanwhile, the wet high-intensity magnetic separation (WHIMS) could produce a magnetic concentrate with TREO recovery of ~80% and enrichment ratio of 5.5 under the following conditions: particle size fraction, −106 + 75 μm; feed flow rate, 3.2 L/min; magnetic induction, 0.8 T. These results indicate that combining DHIMS and WHIMS to upgrade the low-grade REE ore from the Khalzan Buregtei deposit is an effective approach.