Chalcopyrite is the richest copper sulfide mineral in the world, but it is also the most resistant to biohydrometallurgical processing. To promote the bioleaching of chalcopyrite, a nonionic surfactant, t-octyl phenoxy polyethoxy ethanol (Triton X-100), was employed in this paper. Action of Triton X-100 in chalcopyrite leaching using Acidithiobacillus ferrooxidans was explored in shake flasks. Results showed that 30 mg·L⁻¹ of Triton X-100 increased the bioleaching yield of copper by 42.21% compared to the process without additive after 24 days. Under the stress of Triton X-100, the bioleaching efficiency of chalcopyrite slightly dropped at an early stage, but remarkably increased afterwards. XRD and XPS analysis of the leach residues demonstrated that potassium jarosite and elemental sulfur resulted in surface leaching passivation. Surfactant Triton X-100 appeared to induce the oxidation of elemental sulfur by bacteria owing to the increase in the sulfur surface hydrophobicity. These results suggest that Triton X-100 itself has no ability to leach chalcopyrite, but under its induction, the bioleaching of chalcopyrite can be enhanced due to the removal of the passivation layer.

The dispersion behavior of particles is of great significance in selective flocculation flotation. The interfacial interaction between coal and the main impurity mineral (kaolinite) particles with the effect of an anionic polyacrylamide (PAM A401) was explored by the extended Derjagin–Landau–Verwey–Overbeek (DLVO) theory. The involved surface free energy components of fine mineral particles were estimated using the van Oss-Chaudhury-Good theory and Washburn equation. After adsorption of PAM A401, the range and absolute value of the hydrophobic interaction V_HA of the coal particles decreased, the electrostatic repulsive potential increased, and the total potential energy changed from −1.66 × 10⁵ to −4.03 × 10⁴ kT at the separation distance of 5 nm. For interactions between the kaolinite and coal particles after PAM A401 adsorption, the electrostatic repulsive potential increased and the hydrophilic repulsive potential energy decreased. The energy barrier at the separation distance of 0.2 nm decreased from 2.78 × 10⁴ to 2.29 × 10⁴ kT. The total potential energy between the kaolinite and coal particles after PAM A401 adsorption was still repulsive, and the range of the repulsive interaction increased from ~0.05 to 47 nm to ~0.05 to 50 nm. The total potential energy of the coal particles after PAM A401 adsorption was still attractive. This behavior of coal and kaolinite particles with the effect of PAM A401 indicates the possibility of enhanced fine coal separation by the method of selective flocculation flotation.

Gold leaching is an important process to extract gold from ore. Conventional alkaline cyanide process and alternative nontoxic lixiviants including thiosulfate, thiourea, thiocyanate, and halogen have been widely investigated. However, density functional theory (DFT) study on the gold complexes Au(CN)₂⁻, Au(S₂O₃)₂³⁻, Au[SC(NH₂)₂]₂⁺, Au(SCN)₂⁻, and AuCl₂⁻ required for discovering and designing new highly efficient and environmentally friendly gold leaching reagents is lacking, which is expected to support constructive information for the discovery and designation of new high-efficiency and environmentally friendly gold leaching reagents. In this study, the structure information, electron-transferring properties, orbital interaction, and chemical bond composition for complexes Au(CN)₂⁻, Au(S₂O₃)₂³⁻, Au[SC(NH₂)₂]₂⁺, Au(SCN)₂⁻, and AuCl₂⁻ depending on charge decomposition analysis (CDA), natural bond orbital (NBO), natural resonance theory (NRT), electron localization function (ELF), and energy decomposition analysis (EDA) were performed based on DFT calculation. The results indicate that there is not only σ-donation from ligand to Au⁺, but also electron backdonation from Au⁺ to ligands, which strengthens the coordinate bond between them. Compared with Cl⁻, ligands CN⁻, S₂O₃²⁻, SC(NH₂)₂, and SCN⁻ have very large covalent contribution to the coordinate bond with Au⁺, which explains the special stability of Au-CN and Au-S bonds. The degree of covalency and bond energy in Au–ligand bonding decreases from Au(CN)₂⁻, Au(S₂O₃)₂³⁻, Au[SC(NH₂)₂]₂⁺, Au(SCN)₂⁻, to AuCl₂⁻, which interprets the stability of the five complexes: Au(CN)₂⁻ > Au(S₂O₃)₂³⁻ > Au[SC(NH₂)₂]₂⁺ > Au(SCN)₂⁻ > AuCl₂⁻.