The present article aims to develop a technological scheme for processing zinc ferrite residue, which typically forms during the leaching of zinc calcine. This semi-product is currently processed through the Waelz process, the main disadvantage of which is the loss of precious metals with the Waelz clinker. The experimental results of numerous experiments and analyses have verified a technological scheme including the following operations: sulfuric acid leaching of zinc ferrite residue under atmospheric conditions; autoclave purification of the resulting productive solution to obtain hematite; chloride leaching of lead and silver from the insoluble residue, which was produced in the initial operation; and cementation with zinc powder of lead and silver from the chloride solution. Utilizing such an advanced methodology, the degree of zinc leaching is 98.30% at a sulfuric acid concentration of 200 g/L, with a solid-to-liquid ratio of 1:10 and a temperature of 90 °C. Under these conditions, 96.40% Cu and 92.72% Fe form a solution. Trivalent iron in the presence of seeds at a temperature of 200 °C precipitates as hematite. In chloride extraction with 250 g/L NaCl, 1 M HCl, and a temperature of 60 °C, the leaching degree of lead is 96.79%, while that of silver is 84.55%. In the process of cementation with zinc powder, the degree of extraction of lead and silver in the cement precipitate is 98.72% and 97.27%, respectively. When implementing this scheme, approximately 15% of the insoluble residue remains, containing 1.6% Pb and 0.016% Ag.

The present article aims to develop a technological scheme for processing zinc ferrite residue, which typically forms during the leaching of zinc calcine. This semi-product is currently processed through the Waelz process, the main disadvantage of which is the loss of precious metals with the Waelz clinker. The experimental results of numerous experiments and analyses have verified a technological scheme including the following operations: sulfuric acid leaching of zinc ferrite residue under atmospheric conditions; autoclave purification of the resulting productive solution to obtain hematite; chloride leaching of lead and silver from the insoluble residue, which was produced in the initial operation; and cementation with zinc powder of lead and silver from the chloride solution. Utilizing such an advanced methodology, the zinc leaching degree is 98.30 % at a sulfuric acid concentration of 200 g/L, with a solid-to-liquid ratio of 1:10 and a temperature of 90 °C. Under these conditions, 96.40 % Cu and 92.72 % Fe go into solution. Trivalent iron in the presence of nuclei at a temperature of 200 °C precipitates as hematite. In chloride extraction with 250 g/L NaCl, 1M HCl and a temperature of 60 °C, the leaching degree of lead is 96.79 %, while that of silver is 84.55 %. In the process of cementation with zinc powder, the degree of extraction of lead and silver in the cement precipitate is 98.72 % and 97.27 %, respectively. When implementing this scheme, approximately 15 % of the insoluble residue remains, containing 1.6 % Pb and 0.017 % Ag.

Galvanization is a widely employed process in metallurgical practice for the protection of steel products against corrosion. The spent, hydrochloric acid solutions generated from this process contain considerable amounts of valuable metals, such as zinc (40-60 g/L) and manganese (1-2 g/L). The recovery of these metals is not only economically advantageous but also environmentally sustainable. The fundamental difficulty in the treatment of such solutions lies in their high acidity and the high concentration of iron (160-190 g/L). Based on a comprehensive literature review, the following scheme can be proposed for the recovery of zinc in the form of zinc sulfate. The processes include removal of iron through precipitation as iron hydroxide; the precipitation of zinc and manganese hydroxides; dissolution of these precipitates in sulfuric acid to generate a sulfate solution. The objective of the present study is to carry out a detailed thermodynamic evaluation of the possible chemical interactions at each stage using the HSC Chemistry software, version 10.5, specifically the modules: Reaction Equations and Eh-pH Diagrams.

Soda smelting slag is generated during the processing of desulphurized paste from recycling of waste lead acid batteries and other lead containing secondary materials in tilting rotary furnaces, and it represents about 9.5 % of the mass of the charge entering the furnace. This type of slag is very reactive. It contains toxic substances (lead, arsenic and antimony) and because of the presence of alkaline compounds it is corrosive. In addition, it decomposes in contact with air, forming dust. Therefore, such hazardous product cannot be disposed without preliminary stabilization. The aim of the present work is to investigate the possibility for environmentally utilization of this slag. The conducted experiments have shown the possibility of extracting the sodium sulfate from the soda slag only by grinding in an aqueous medium with subsequent filtration. The resulting solution is subjected to sulfide purification from heavy non-ferrous metals and subsequent crystallization for obtaining a commercial product sodium sulfate. These processes could be easily implemented in companies that have applied Engitec’s technology for recycling spent lead-acid batteries and production of sodium sulfate from desulfurization of the lead paste. Implementation of these processes will eliminate the need for slag stabilization and prevent environmental problems.

This study investigates the feasibility of efficiently recovering lead and silver from a lead cake by applying a combined process of chlorination roasting followed by acid leaching. The lead cake is obtained after the sulfuric acid leaching of zinc ferrite residues generated during the hydrometallurgical treatment of zinc calcine. The influence of roasting temperature, the mass ratio between lead cake and NaCl, and the roasting duration on metal extraction was systematically investigated to determine optimal process conditions. The most efficient parameters were identified as roasting at 550 °C for 1.5 hours with a lead cake to NaCl ratio of 1:3, followed by leaching the roasted product in 1 M HCl. Under these conditions, the remaining Pb and Ag contents in the final solid residue were reduced to 0.90 % and 0.0027 %, respectively, indicating nearly complete chlorination and subsequent dissolution of both metals. A comparative evaluation showed that the combined chlorination roasting–leaching approach resulted in higher recovery rates (Pb 98.67 %, Ag 98.09 %) and a smaller final residue mass than direct chloride leaching (Pb 96.79 %, Ag 84.55 %). Therefore, the proposed method is demonstrated to be more effective and environmentally advantageous for recovering valuable metals from industrial waste materials.

This study evaluates an integrated approach for recovering lead and silver from lead cake through chlorination roasting followed by acid leaching. The lead cake originates from sulfuric acid leaching of zinc ferrite residues obtained during the hydrometallurgical processing of zinc calcine. The effects of roasting temperature, lead cake-to-NaCl mass ratio, and roasting duration on metal recovery were systematically examined to determine optimal process conditions. Based on the experimental results, roasting at 550 °C for 1.5 h with a lead cake-to-NaCl mass ratio of 1:3, followed by leaching in 1 M HCl, was selected as a representative and sufficiently effective condition for the combined process. Under these conditions, nearly complete dissolution of Pb and Ag was achieved, reducing their contents in the final solid residue to 0.90% and 0.0027%, respectively. Compared to direct chloride leaching, the combined process provided higher extraction efficiencies (Pb 98.67%, Ag 98.09%) and a lower final residue mass (34% vs. 45%). The roasting step enables the solid-state conversion of PbSO₄ into highly soluble chloride phases (PbCl₂ and Pb(OH)Cl), while ZnFe₂O₄, Fe₂O₃ and SiO₂ remain stable and form the inert matrix of the residue. Acid leaching at a lower solid-to-liquid ratio (1:10) ensures near-complete dissolution of Pb and Ag, whereas aqueous leaching at a high ratio (1:100) results in incomplete Pb removal. The compliance leaching test (EN 12457-2) confirmed that the residue produced after the optimized two-step treatment meets the EU criteria for inert waste. Overall, the proposed combined process enhances Pb and Ag recovery, minimizes environmental risk, and offers a technically robust and sustainable route for treating lead-containing industrial residues.