Bauxite residue, known as red mud, is a by-product of alumina production using the Bayer process. Currently, its total global storage amounts to over 4.6 billion tons, including about 600 million tons in Russia. The total global storage of red mud occupies large areas, leading to environmental damage and increasing environmental risks. Moreover, it contains a significant amount of sodium, which is easily soluble in subsoil water; therefore, a sustainable approach for comprehensive recycling of red mud is necessary. The bauxite residue contains valuable elements, such as aluminum, titanium, and scandium, which can be recovered using liquid media. In recent years, many methods of recovery of these elements from this waste have been proposed. This paper provides a critical review of hydrometallurgical, solvometallurgical, and complex methods for the recovery of valuable components from red mud, namely, aluminum, titanium, sodium, and rare and rare-earth elements. These methods include leaching using alkaline or acid solutions, ionic liquids, and biological organisms, in addition to red mud leaching solutions by extraction and sorption methods. Advantages and disadvantages of these processes in terms of their environmental impact are discussed.

Red mud is an iron-containing waste of alumina production with high alkalinity. A promising approach for its recycling is solid-phase carbothermic roasting in the presence of special additives followed by magnetic separation. The crucial factor of the separation of the obtained iron metallic particles from gangue is sufficiently large iron grains. This study focuses on the influence of Na2SO4 addition on iron grain growth during carbothermic roasting of two red mud samples with different (CaO + MgO)/(SiO2 + Al2O3) ratio of 0.46 and 1.21, respectively. Iron phase distribution in the red mud and roasted samples were investigated in detail by Mössbauer spectroscopy method. Based on thermodynamic calculations and results of multifactorial experiments, the optimal conditions for the roasting of the red mud samples with (CaO + MgO)/(SiO2 + Al2O3) ratio of 0.46 and 1.21 were duration of 180 min with the addition of 13.65% Na2SO4 at 1150 °C and 1350 °C followed by magnetic separation that led to 97% and 83.91% of iron recovery, as well as 51.6% and 83.7% of iron grade, respectively. The mechanism of sodium sulfate effect on iron grain growth was proposed. The results pointed out that Na2SO4 addition is unfavorable for the red mud carbothermic roasting compared with other alkaline sulfur-free additives.

Red mud is a by-product of alumina production from bauxite ore by the Bayer method, which contains considerable amounts of valuable components such as iron, aluminum, titanium, and scandium. In this study, an approach was applied to extract iron, i.e., carbothermic reduction roasting of red mud with sodium and potassium carbonates followed by magnetic separation. The thermodynamic analysis of iron and iron-free components’ behavior during carbothermic reduction was carried out by HSC Chemistry 9.98 (Outotec, Pori, Finland) and FactSage 7.1 (Thermfact, Montreal, Canada; GTT-Technologies, Herzogenrath, Germany) software. The effects of the alkaline carbonates’ addition, as well as duration and temperature of roasting on the iron metallization degree, iron grains’ size, and magnetic separation process were investigated experimentally. The best conditions for the reduction roasting were found to be as follows: 22.01% of K2CO3 addition, 1250 °C, and 180 min of duration. As a generalization of the obtained data, the mechanism of alkaline carbonates’ influence on iron grain growth was proposed.

Red mud is a hazardous waste of the alumina industry that contains high amounts of iron, aluminum, titanium and rare-earth elements (REEs). One of the promising methods for the extraction of iron from red mud is carbothermic reduction with the addition of sodium salts. This research focuses on the process of hydrochloric high-pressure acid leaching using 10 to 20% HCl of two samples of non-magnetic tailings obtained by 60 min carbothermic roasting of red mud at 1300 °C and the mixture of 84.6 wt.% of red mud and 15.4 wt.% Na2SO4 at 1150 °C, respectively, with subsequent magnetic separation of metallic iron. The influence of temperature, leaching duration, solid-to-liquid-ratio and acid concentration on the dissolution behavior of Al, Ti, Mg, Ca, Si, Fe, Na, La, Ce, Pr, Nd, Sc, Zr was studied. Based on the investigation of the obtained residues, a mechanism for passing valuable elements into the solution was proposed. It has shown that 90% Al, 91% Sc and above 80% of other REEs can be dissolved under optimal conditions; Ti can be extracted into the solution or the residue depending on the leaching temperature and acid concentration. Based on the research results, novel flowsheets for red mud treatment were developed.

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Red mud is a by-product of alumina industry which is currently almost completely stored in landfill sites without further use. It contains considerable amounts of valuable components such as iron, aluminum, titanium and rare-earth metals. The reduction smelting of red mud was carried out in laboratory scale to recover iron and obtain slag suitable for use in the construction industry. It has been shown that it is expedient to obtain pig iron and slag from the unprocessed red mud. Those two are suitable for the subsequent leaching of aluminum, titanium and rare-earth metals. It is practical to process dealkalized red mud, with composition adjustment by CaO and Al2O3 addition, in order to obtain pig iron and slag in the form of aluminous clinker.

Zinc plant residue (ZPR) is a secondary material generated during hydrometallurgical zinc production that contains considerable contents of valuable elements such as Zn, Cu, Fe, Pb, Cd, Ag, In, Ga, Tl. Zinc, copper and accompanying elements in ZPR are in different minerals, mainly in the ferrites. A promising approach for recycling ZPR is the sulfating roasting using iron sulfates followed by water leaching. In this study, the composition of ZPR and the obtained products were thoroughly investigated by various methods including X-ray diffraction analysis (XRD), chemical phase analysis and Mössbauer spectroscopy. The effect of temperature, amount of iron sulfates and roasting time on the conversion of valuable metals into a water-soluble form was thermodynamically and experimentally studied both using pure ferrites and ZPR. Based on the results of time-resolved XRD analysis and synchronous thermal analysis (STA), a mechanism of the sulfation roasting was elucidated. The rate-controlling step of zinc and copper sulfation process during the ZPR roasting was estimated. The sulfating roasting at 600 °C during 180 min with the optimal Fe2(SO4)3∙9H2O addition followed by water leaching enables to recover 99% Zn and 80.3% Cu, while Fe, Pb, Ag, In, Ga retained almost fully in the residue.

Изучены научные основы пирометаллургической технологии переработки пылей электросталеплавильного производства, содержащих ферриты цинка. Выполнен термодинамический анализ разложения феррита цинка известью. Анализ полученных расчетных данных показал, что для разложения более чем 90 % ZnFe 2 O 4 необходимо добавить к пыли не менее 46 % CaO, а для разложения более чем 95 % ZnFe 2 O 4-не менее 60 % CaO. Результаты расчетов проверены экспериментально на лабораторной печи. Экспериментальная прокалка пыли на воздухе с добавлением извести в количестве 60 % от массы пыли при температуре 1000 °C с временем выдержки 4 ч подтвердила, что процесс разложения феррита цинка оксидом кальция с образованием оксида цинка и двухкальциевого феррита имеет место. При этом также были получены возгоны в количестве 50 кг на 1 т пыли, содержащие 29 % свинца и 15 % цинка. Процесс прокалки пыли с известью можно применять для перевода цинка из феррита в растворимую оксидную форму. В результате прокалки могут быть получены промежуточные продукты для извлечения цинка и свинца. После выщелачивания цинка возможно получение железосодержащего продукта, который может быть востребован в черной металлургии. Использованный подход имеет ряд технологических преимуществ по сравнению с известной технологией вельцевания цинксодержащих металлургических пылей. В частности, процесс протекает при более низкой температуре (1000 °C) по сравнению с известной технологией (1250 °C), исключается вторая стадия вельцевания, необходимая для очистки от галогенидов поступающего на выщелачивание оксида цинка, значительно сокращается расход кокса, а также упрощается очистка газов от пыли из-за уменьшения количества возгонов в 6-8 раз.