Iron control in the atmospheric acid leaching (AL) of nickel laterite was evaluated in this study. The aim was to decrease acid consumption and iron dissolution by iron precipitation during nickel leaching. The combined acid leaching and iron precipitation process involves direct acid leaching of the limonite type of laterite followed by a simultaneous iron precipitation and nickel leaching step. Iron precipitation as jarosite is carried out by using nickel containing silicate laterite for neutralization. Acid is generated in the jarosite precipitation reaction, and it dissolves nickel and other metals like magnesium from the silicate laterite. Leaching tests were carried out using three laterite samples from the Agios Ioannis, Evia Island, and Kastoria mines in Greece. Relatively low acid consumption was achieved during the combined precipitation and acid leaching tests. The acid consumption was approximately 0.4 kg acid per kg laterite, whereas the acid consumption in direct acid leaching of the same laterite samples was approximately 0.6–0.8 kg acid per kg laterite. Iron dissolution was only 1.5–3% during the combined precipitation and acid leaching tests, whereas in direct acid leaching it was 15–30% with the Agios Ioannis and Evia Island samples and 80% with the Kastoria sample.

Acid bioleaching of a low-grade and polymetallic sulfide concentrate was studied, in order to determine suitable feed material particle size and pH for efficient leaching of valuable metals. The sulfide concentrate consisted of pyrite (50 wt %), pyrrhotite (31 wt %), quartz (10 wt %) and lower amounts of cobalt, nickel, zinc and copper (each <1 wt %). After adaptation of microorganisms in shake flasks, stirred tank tests were conducted at different pH levels and supplementing feed material at different particle sizes (milled to d80 < 150 µm, <50 µm, <28 µm, <19 µm). The operation at pH 1.8 was seen prone to iron precipitation, while this was not observed at a pH between 1.3 and 1.5. Additional milling to decrease particle size from the initial d80 < 150 µm had a major positive effect on cobalt- and nickel-leaching yields, proposing that at least d80 < 28 µm should be targeted. The best leaching yields for the main economic elements, cobalt and nickel, were 98% and 94%, respectively, reached with d80 < 19 µm at pH 1.3. However, it was considered that at pH 1.5, similar results could be obtained. This research sets the basis for continuing the experiments at a continuous pilot scale.

Copper slags produced in vast quantities in smelting operations could be considered as secondary material sources instead of stockpiling them in landfills. This study investigates the recovery of valuable metals from copper slag and the valorisation of the leach residue as construction material in line with the principles of a circular economy. By taking into account that the environmental characterization of the as-received copper slag did not allow its disposal in landfills without prior treatment, chemical and biological leaching were tested for the recovery of metals. Pre-treatment with acids, namely HNO₃ and H₂SO₄, resulted in the extraction of several target metals and the production of an almost inert waste. Despite the clearly better oxidative conditions prevailing in the bioleaching reactors, chemical leaching resulted in the higher dissolution of Cu (71% vs. 51%), Co (70% vs. 36%), and Zn (65% vs. 44%). The acid consumption was much lower during the bioleaching experiments compared to the chemical leaching. The bioleach residue was suitable for its use as supplementary cementitious material, showing a better performance than the reference sample without causing any detrimental effects to the calcium aluminate cement (CAC) quality. The complete valorisation of copper slags is expected to improve the economics of the process, by avoiding landfill costs and producing saleable products with high added value.

Sulfate-rich mine water must be treated before it is released into natural water bodies. We tested ethanol as substrate in bioreactors designed for biological sulfate removal from mine water containing up to 9 g L⁻¹ sulfate, using granular sludge from an industrial waste water treatment plant as inoculum. The pH, redox potential, and sulfate and sulfide concentrations were measured twice a week over a maximum of 171 days. The microbial communities in the bioreactors were characterized by qPCR and high throughput amplicon sequencing. The pH in the bioreactors fluctuated between 5.0 and 7.7 with the highest amount of up to 50% sulfate removed measured around pH 6. Dissimilatory sulfate reducing bacteria (SRB) constituted only between 1% and 15% of the bacterial communities. Predicted bacterial metagenomes indicated a high prevalence of assimilatory sulfate reduction proceeding to formation of l-cystein and acetate, assimilatory and dissimilatory nitrate reduction, denitrification, and oxidation of ethanol to acetaldehyde with further conversion to ethanolamine, but not to acetate. Despite efforts to maintain optimal conditions for biological sulfate reduction in the bioreactors, only a small part of the microorganisms were SRB. The microbial communities were highly diverse, containing bacteria, archaea, and fungi, all of which affected the overall microbial processes in the bioreactors. While it is important to monitor specific physicochemical parameters in bioreactors, molecular assessment of the microbial communities may serve as a tool to identify biological factors affecting bioreactor functions and to optimize physicochemical attributes for ideal bioreactor performance.

Municipal solid waste incineration bottom ash (MSWI BA) is the main output of the municipal solid waste incineration process, both in mass and volume. It contains some heavy metals that possess market value, but may also limit the utilization of the material. This study illustrates a robust and simple heap leaching method for recovering zinc and copper from MSWI BA. Moreover, the effect of autotrophic and acidophilic bioleaching microorganisms in the system was studied. Leaching yields for zinc and copper varied between 18–53% and 6–44%, respectively. For intensified copper dissolution, aeration and possibly iron oxidizing bacteria caused clear benefits. The MSWI BA was challenging to treat. The main components, iron and aluminum, dissolved easily and unwantedly, decreasing the quality of pregnant leach solution. Moreover, the physical nature and the extreme heterogeneity of the material caused operative requirements for the heap leaching. Nevertheless, with optimized parameters, heap leaching may offer a proper solution for MSWI BA treatment.

Phosphorus and uranium are both vital elements for society. In recent decades, fears have arisen about the future availability of low-cost phosphorus and uranium. This has resulted in pressure to de-centralize production of both elements by utilizing lower-grade or complex deposits. The research presented here focused on phosphorus-containing apatite ores with uranium impurities; in order to separate uranium by selective and sequential bioleaching before phosphorus leaching. This would create an alternative process route for solvent-extraction, used to remove/recover uranium from the phosphorus acid product of apatite H₂SO₄ wet process. In this work, it was seen that the used fluorapatite ore required 24 h leaching at pH 1 by H₂SO₄ to result in 100% leaching yield for phosphorus. As this ore did not contain much uranium, an artificial fluorapatite-uranium ore was prepared by mixing standard uranium ore and fluorapatite. The research with this ore showed that 89% of uranium dissolved in 3 days at pH > 2 and leaching was improved by applying Fe³⁺ oxidant. In these conditions only 4% of phosphorus was leached. By prolonged (28 days) leaching 95% uranium yield was reached. According to the experiments, the iron in the uranium leach solution would be mainly Fe³⁺, which allows the use of H₂O₂ for uranium recovery and then direct use of spent leachate for another uranium leaching cycle. After the dissolution of uranium, 90% of phosphorus was dissolved by decreasing the pH to 1.3. This was done by bioleaching, by utilizing biogenic sulfur oxidation to sulfuric acid.