Red mud is a solid waste generated during the extraction of alumina from bauxite ores by the Bayer method [1
]. At the present time, the main treatment methods for red mud include stockpiling by damming, direct sea-fill, and sea-fill after neutralization [2
]. Globally, about 4 billion tons of red mud are currently stored [3
]. The lack of efficient processing technologies is the main reason why red mud is out of use and accumulated in special sludge storage facilities, which have an adverse impact on the environment and have even led to technogenic catastrophe [3
]. It has found that red mud is proved to be a promising material for obtaining coagulants and can be used as an adsorbent for heavy metals in wastewater [5
]; it can also be utilized in a wide range of catalytic applications [6
], as well as for a production of various building materials [7
The typical contents of valuable components in red mud such as Ti, Si, Fe, Na and Al are 2 to 12%, 1 to 9%, 14 to 45%, 1 to 6% and 5 to 14%, respectively [8
]. Moreover, red mud also contains a small proportion of REEs in the range of 0.05 to 0.17% [9
]. It is important to note that the Sc content in red mud is in the range of 0.013 to 0.039% [10
], which is quite significant. The actual composition of red mud depends on the bauxite mineralogy and different technological parameters of the Bayer process [11
]. The iron content in red mud is comparable with poor iron ore [12
], which is an important driver for research on the processing of red mud.
Many processes have been proposed for the recovery of valuable components from red mud [9
]. In the last years, there has been a high interest for the selective recovery of scandium [10
]. Direct leaching processes using different lixiviants such as alkaline solutions [17
], hydrochloric and sulfuric acids [19
], organic acids [26
] and ionic liquids [28
] have been thoroughly investigated. However, direct leaching methods are either non-selective or ineffective for the extraction of titanium and REEs.
The high content of iron in red mud is the main hindering factor for the extraction of REEs from the solutions obtained by red mud acid leaching [29
]. For this reason, the most promising way to comprehensively recycle red mud is by using combined pyro- and hydrometallurgical methods with the recovery of iron by pyrometallurgical methods followed by the leaching of valuable components from the iron-depleted red mud [31
]. The main pyrometallurgical ways for iron recovery from red mud are reduction smelting and carbothermic roasting with the subsequent magnetic separation of reduced iron. The reduction smelting either requires a high temperature and uses different fluxes such as CaO [32
], or needs the preliminary extraction of alumina [36
] for the obtaining of slags with a low viscosity and melting temperature. The more favorable way is carbothermic roasting followed by magnetic separation, which is favorable due to its lower energy consumption compared with reduction smelting. Many studies [37
] have shown that direct carbothermic roasting of red mud with different additions followed by magnetic separation results in the obtaining of iron concentrates with about 95% of iron content at temperatures below 1200 °C.
There are several studies describing the leaching of valuable elements from tailings obtained by the carbothermic roasting of red mud with alkaline salts followed by magnetic separation. Such a two-step process followed by alumina extraction using alkaline leaching was proposed for red mud treatment [43
], but alkaline leaching of the tailings led to the recovery of only alumina and sodium without the extraction of other valuable elements. It has shown [45
] that the addition of sodium salts such as Na2
significantly improves the recovery of Al, Fe, Ti and Si from the tailings by sulfuric acid leaching at 30 °C.
The dissolution of silicon from red mud during an acid leaching can lead to the formation of silica gel, which causes significant difficulties for the filtration of the leached solution [46
], so different methods were proposed to avoid it. The comparative atmosphere leaching of the tailings by different inorganic acids, namely, hydrochloric, nitric, sulfuric, and phosphoric [47
] has indicated that only phosphoric acid allows the selective removal of SiO2
with substantial enrichment of Sc2
in the residue. A suggested two-step process includes SiO2
extraction by phosphoric acid leaching followed by Al2
extraction from the residue by the leaching using sodium hydroxide. The other approaches, which allows silica gel formation to be avoided, are high-pressure acid leaching (HPAL) or using a high liquid-to-solid ratio [48
]. The comparative study of the dissolution efficiency of the slags after red mud reduction smelting at atmosphere leaching and high-pressure leaching using HCl and H2
have shown that high-pressure hydrochloric acid leaching allows for the extraction of more than 90% Al and above 95% Y, La, and Nd, as well as up to 80% Sc with the co-dissolution of Si and Ti below 5% [30
]. The use of hydrochloric acid for HPAL for the aluminum extraction in different high-silica materials, e.g., coal ash [49
], and boehmite-kaolinite bauxites [50
], resulted in higher aluminum extraction compared with other mineral acids [52
Thus, a promising way for the extraction of valuable components from red mud is the reduction roasting with obtaining of magnetic iron concentrate and non-magnetic tailings enriched in aluminum, titanium and REEs followed by the treatment of the tailings by high-pressure hydrochloric acid leaching.
In this paper, we study hydrochloric acid leaching of two different kinds of the tailings after the carbothermic roasting and magnetic separation of red mud. The leaching behavior of major elements (Al, Ti, Ca, Si, Fe, Na, Mg) and minor elements (Sc, Y, La, Ce, Nd, Pr, Zr) under atmosphere and high-pressure conditions were investigated. The most influencing factors on the leaching process were identified, and optimal leaching conditions were determined. The obtained solutions and the solid residues were characterized by different methods, so a leaching mechanism to extract valuable components was proposed. Two flowsheets for the processing of the both non-magnetic tailings were developed.
The HPAL of the tailings using the diluted hydrochloric acid allows the selective separation of REEs and Al from Ti and SiO2
with high effectivity. The obtained residue can contain more than 20% TiO2
and over 50% SiO2
) which indicates that it is a valuable raw material for Ti extraction. To extract Ti, various leaching treatments have been suggested, such as using the mixture of hydrogen peroxide and sulfuric acid with a suppression of Si dissolution [71
], as well as using alkaline solutions [72
], phosphoric acid [47
], etc. However, the most promising way for Ti extraction from the residue is the leaching by concentrated hydrochloric acid with its dissolution and subsequent precipitation of TiO2
]. The obtained amorphous SiO2
can be used for the production of white carbon black [74
REEs can be relatively simply recovered from the leached solutions by well-known methods, e.g., solvent extraction, ion-exchange sorption, selective precipitation, etc. [75
]. The remaining aluminum chloride solution can be used as a coagulant for water treatment [62
] or as a raw material for the production of metallurgical alumina [77
]. Thus, a two-step hydrochloric acid leaching process can be implemented.
As reflected by Figure 13
, the application of 20% HCl with S:L ratio of 1:16.5 led to a considerable rise of Al, Ti, Sc extraction degree that enables the consideration of the simultaneous dissolution of these elements with a following stepwise extraction of REEs and titanium from the leached solution. Moreover, after the REE extraction, Al-Ti-containing solution can be regarded as a complex coagulant, which has an increased efficiency for waste water treatment compared with the Ti-free solutions [78
]. However, it should be noted that the use of an excess of acid and liquid-to-solid ratio during HPAL deteriorates the cost efficiency of the process, so a feasibility study is necessary.
The application of the HPAL method using HCl led to a similar Al recovery for the WAT and SSAT samples, but the treatment of the SSAT resulted in a rather lower Sc recovery and a slightly higher acid consumption. Furthermore, the residue obtained by the leaching of the SSAT sample can contain a substantial percentage of elemental sulfur and sodium chloride, so additional stages for their removal is necessary. The process of elemental sulfur precipitation is given in Section 3.2.2
; sodium chloride was likely precipitated during the filtration of the mother liquors. To remove NaCl, washing by water can be applied, and to extract elemental sulfur, distillation [80
] or flotation [81
] can be used. Despite the mentioned disadvantages of the SSAT leaching, the total cost of red mud processing using sodium sulfate may be lower owing to the reduced roasting temperature and improved grindability of the roasted sample. Moreover, the cost of both flowsheets can be additionally reduced by the replacement of a carbonaceous reductant with blast furnace sludge [82
]. In any case, a comparative economic assessment of the treatment of the WAT and SSAT samples is required.
shows two principal flowsheets of red mud processing developed as a result of the study.
The suggested flowsheets aimed at a comprehensive utilization of red mud with the extraction of all valuable components. The recovery of iron by reduction roasting at the first stage at 1150 to 1300 °C allows the reduction of energy costs compared with a reduction smelting with the application of a low-grade coal or carbon-containing waste as a reducing agent.
As mentioned above, hydrochloric acid leaching is an effective method for the alumina recovery, and it was used on a pilot scale to extract Al2
from coal fly ash in China [83
]. To obtain polyaluminum chloride, 30 to 32% HCl is commonly used to dissolve aluminum hydroxide. The process occurs using high-pressure leaching in reactors at 150 to 180 °C depending on the charge material. The use of 3–6 M HCl is due to the possibility of acid regeneration during the thermal hydrolysis of aluminum chloride hexahydrate to obtain alumina. The hydrochloric acid flowsheet involves obtaining alumina by the precipitation of salt from an acidic aluminum chloride solution followed by calcination of the salt to form aluminum oxide. Hydrochloric acid regenerates simply—HCl vapor bubbles through the water, so up to 6 M HCl can be obtained. Hence, the proposed flowsheets can be characterized by a low environmental impact.
It also should be noted that there is a high flexibility of the suggested flowsheets. The process can be organized to obtain high-demand products at the moment. For example, at high prices for alumina, it can be extracted from the solution; otherwise, the coagulants can be obtained. In the same way, based on market conditions, titanium can be passed into solution to obtain complex coagulants, or it can be extracted into a particular product. An engineering and economic evaluation of the flowsheets is possible after obtaining and characterizing the final products, namely, the REE and titanium concentrates, white carbon black, the coagulants, and alumina. However, the above-mentioned advantages and a possibility of obtaining various high-demand products indicate the economic feasibility of the proposed methods.