This paper focuses on the environmental leaching of antimony, a critical mineral, using deep eutectic solvents. Mining residues often contain embedded antimony, posing environmental risks. Deep eutectic solvents, known for being low in toxicity, cost-effective, and environmentally friendly, present a promising avenue for sustainable antimony extraction. The study focuses on optimizing the leaching process through experimental analysis by considering variables such as temperature, time, and percentage of solids. Different deep eutectic solvent (DES) compositions are being studied, including choline chloride with malonic acid, thiourea, and ethylene glycol in different molar ratios, to identify the most effective solvent system for antimony extraction. A sample, originally obtained from mining waste produced via the flotation of antimonite ore, was used to test these three types of DESs. By optimizing the leaching process by changing the ratio of solid and liquid components, as well as the amount of oxidizing agent up to 3 g, iodine, yields of up to 100% were achieved after leaching for 4 h at 100 °C. The aim of the study is to advance sustainable resource management by providing knowledge on an ecological and feasible method of extracting antimony from mine waste, leading to more conscious and efficient resource practices in the mining sector.

Slag is one of the by-products of the energy industry, which is suitable for secondary industrial processing. Although slag has been successfully used in industrial production for several decades, its use does not achieve the level of its potential. Today, to achieve a sustainable construction industry, alternative types of cement have been extensively investigated. Geopolymer is a kind of material which is obtained from the alkaline activator and it can be produced from the industrial wastes or by-products. In this study, SiO₂/Na₂O ratio and the amount of Na₂O in activation solution parameters of alkali-activated materials were tested how they affect the strengths of hardened geopolymers from ground granulated blast furnace slag (GGBFS). Compressive and flexural strength tests were conducted, and the results were analyzed by analysis of variance (ANOVA). Strengths were tested after 7, 28, and 90 days.

This study describes the bioleaching of a flotation concentrate by acidophilic iron- and sulfur-oxidizing bacteria at ambient temperature in shake flasks. The concentrate was prepared from polymetallic Ag–Cu–Sb–siderite ores from Rožňava (Slovakia) and consisted of tetrahedrite and siderite, with small amounts of chalcopyrite, pyrite, and quartz. Biologically driven Cu and S extraction was significantly greater than the extraction in the abiotic control. The oxidation of 1 mol of tetrahedrite (nominally Cu ₁₀ ⁺ Cu ₂ ²⁺ Sb ₄ ³⁺ S ₁₃ ²⁻ ) to produce sulfate involved the transfer of 122 mol of electrons. The stoichiometry of the reaction was confirmed by gas analyses and the concentration profiles of the oxidation products and can be described by the equation Cu ₁₂ Sb ₄ S ₁₃ + 122Fe ³⁺ + 52H ₂ O → 12Cu ²⁺ + 4Sb ⁵⁺ + 13SO ₄ ²⁻ + 122Fe ²⁺ + 104H ⁺ . The rate of bacterial ferrous iron oxidation highly exceeded that of the chemical reduction of Fe ³⁺ at the tetrahedrite surface, as indicated by the high Fe ³⁺ -to-Fe ²⁺ ratio in the leaching liquors. The sulfur entity (S ²⁻ ) of tetrahedrite was oxidized ultimately to sulfate. After 120 days of tetrahedrite bioleaching, the concentration of Cu in the leaching liquor exceeded 4 g L ⁻¹ , which is equivalent to > 80% metal yield. While the recovery of Cu and S increased proportionally with time, the Sb concentration approached 40 mg L ⁻¹ , and subsequently decreased, which indicates the formation of secondary antimony minerals. Electron microprobe analyses of the solid residues showed that the alteration products are distinctly depleted in Cu and S and enriched with Sb and Fe when compared to the primary tetrahedrite. Graphical Abstract

This study demonstrated the successful recovery of zinc, lead, and copper collective concentrates from historical metal-bearing mine tailings (sulfide–polymetallic ore with a composition of 7.38% Zn, 1.45% Pb, and 0.49% Cu) using froth flotation techniques, which were originally developed during uranium ore mining. Froth flotation techniques were used to justify suitability for recovering metals. The effects of a dosage of the foaming agent Polyethylene glycol (PEG 600) at 50 and 100 g t⁻¹, collector types Aerophine 3418A (AERO), Danafloat 067 (DF), and potassium ethyl xanthate (KEX) at 50 and 80 g t⁻¹, and a suspension density of 300 and 500 g L⁻¹ on froth flotation collective concentrates were investigated. The final collective concentrate achieved recoveries exceeding 91% for lead (Pb), 88% for copper (Cu), and 87% for zinc (Zn). The obtained concentrates were analyzed using Atomic Absorption Spectroscopy (AAS) and X-ray Fluorescence Spectrometry (XRF), while selected samples were further examined via Scanning Electron Microscopy (SEM) with Energy Dispersive X-ray Spectroscopy (EDS). The resulting sulfide concentrates can subsequently be treated using suitable hydrometallurgical techniques. The application of these concentrates in metal production would help reduce the environmental burden of mining activities.

The history of exploitation of gold from the Danube River’s sandy gravels is centuries long. The extraction of valuable heavy minerals (VHM) concentrate was never intended. Our aim is to find out an effective separation process to produce monomineral concentrates of the following minerals: garnet, ilmenite, zircon, monazite, magnetite, rutile, gold. The essential condition is to use no chemistry (no flotation, leaching, activating). The experimental concentrates were prepared by sluicing on active river channel. Next, the separation results were achieved using gravity and electromagnetic methods with different magnetic intensities. The prepared rutile contained from 63.3% TiO₂ to 87% TiO₂. The ilmenite concentrate contained 20.5% TiO₂ and 39.2% ilmenite. The garnet concentrate contained 94% garnet. The monazite concentrate contained 86.6% monazite, and the sum of REE oxides was 50.1%. The zircon concentrate containing 63.7% ZrO₂ means that the prepared concentrate contained 96.1% zircon.