The case of low-level Sc leaching with mineral acid

A different approach to Sc leaching from BR was recently published under the SCALE project (Balomenos et al, 2021a). The leaching conditions were selected based on criteria of cost and effectiveness of the produced PLS in a subsequent ion-exchange process to extract Sc into a marketable concentrate. To achieve low cost in leaching, sulfuric acid was used at high pulp densities and to achieve effectiveness of the ion-exchange process iron and titanium in the PLS were kept below 200 mg/l and 50 mg/l respectively. The conditions meeting these demands were leaching the BR with 0.26 – 0.28 kg conc. sulfuric acid/kg BR at 85oC for 1 h with 40% solid to liquid ratio. At these conditions silica gel formation was avoided (Si concentration in PLS 120-150 mg/l). The process was demonstrated at pilot scale at MYTILINEOS in Greece, where on average from 1 t BR, 17,15g Sc were dissolved into PLS, consuming 0.27t of H2SO4 (21% Sc leaching yield, 12-14 mg/l of Sc in the PLS). With more than 10t of BR leached in this way, sufficient PLS was generated to process in ion-exchange pilot utilizing II-VI’s SIR resin, resulting in a solid concentrate with 22% wt Sc (Balomenos et al, 2021b).

Sulfuric acid baking

The BR baking process with H2SO4 was developed, after extensive research, as a suitable method to increase the selective solubilization of REEs against major metals in BR, reducing the consumption of large quantities of chemical reagents (Borra et al., 2016c; Liu et al., 2017; Narayanan et al., 2018; Onghena et al., 2017). In this process, BR is digested at 100-120 ° C with concentrated H2SO4 to form Fe, Al, Ti and REE sulphates, and baking at 650-750 ° C to break down non-stable sulfates such as Fe, Al, Ti in the form of oxides. REE’s sulfate salts remain in the form of sulfates even at 750 ° C allowing their selective solubilization in a subsequent step using water. The solid residue from the process contains mainly oxides Al, Ti, Fe, Si and CaSO4. This process results in the selective extraction of REEs with 60% Sc-90% REE recoveries (Borra et al., 2016c). In a similar study, 75% Sc-88% REE recoveries have been reported (Narayanan et al., 2018). The factors influencing the selective extraction are the amount of acid in the BR, the baking temperature and the baking time. A big advantage of the method is the reduction in acid consumption and the high selectivity of REE over Fe, Al, Si, Ti. The slow rate of REEs water leaching after the sulfate baking stage, which is reported between 2-7 days, and the SO2 emissions are the main drawbacks of the process.

Combined pyro-hydrometallurgical leaching techniques

The pyrometallurgical recovery of Fe from BR is a stepping stone for the development of holistic exploitation, valorizing about half the amount of BR and adjusting the properties of the slag to be further processed for critical metals extraction or to be used as building material. Iron recovery through the pyro-metallurgical BR treatment can be achieved by reductive smelting or by reductive roasting and magnetic separation; REE oxides cannot be reduced by carbon and therefore will never follow the metal phase produced in such processes and will be found in enhanced concentrations in the produced slag or from the non-magnetic fraction respectively.
Reductive smelting processes can be applied by several technologies (Corex, Finex, Hismelt, Romelt, AusIron and Electic Arc Furnace EAF) to BR for the production of pig iron (Panov et al., 2012; Xenidis et al., 2016). So far, two methods have been applied on a large scale for the BR reductive smelting: the Romelt method (Mishra and Bagchi, 2002) and the electric arc furnace (EAF) (Balomenos et al., 2011; Grzymek et al., 1982; Guccione, 1971; Logomerac, 1979). The Moscow Institute for Steel and Alloys (MISA), with National Aluminium Company Limited (NALCO) and Romelt-SAIL (India) Ltd., studied the pyro metallurgical processing of BR using the Romelt method (Mishra and Bagchi, 2002). The advantage of this method is that it accepts feed materials with moisture levels of up to 10% wt. The main disadvantage is the high energy consumption and the production of poor quality of pig iron with a high concentration of S and P (Panov et al., 2012). In the EAF reductive smelting, a mixture of BR, carbon and fluxes, at 1500-1700 °C to form pig iron with> 95% iron recovery (Balomenos et al., 2014; Balomenos et al., 2011; Balomenos et al., 2016). Recovery of residual iron can be further improved by later magnetic separation of slag dust (Borra et al., 2016b; Ercas and Apak, 1997). Post-melting slag can be used to produce rockwool or building materials (Balomenos et al., 2016; Raspopov et al., 2013) and for the recovery of non-ferrous metals and REE (Alkan et al., 2017 Borra et al., 2016b; Udy, 1958). REE is extracted from the slag using strong inorganic acids (Sargic and Logomerac, 1974; Shaoquan and Suqing, 1996). In particular, slag extraction with 60% wt. H2SO4 dissolves virtually all of the Al and Ti contents residing predominantly in the silicon oxide (Udy, 1958), while extraction with inorganic acids HCl and HNO3results in high REE recovery yields (60-100%) in solids to liquid (1/50) at 90 ° C for 1 hour at 2-3 mol/L acid concentration (Borra et al., 2016b). The use of H2SO4 in the same conditions showed similar recovery for Sc but lower for the remaining rare earths probably due to the formation of calcium sulphate. It is clear that a high acid concentration is required to dissolve REE from the slag, however, the number of studies presented in the literature is limited in relation to the extraction of slag for Sc recovery using sulfuric acid as well as the description of extraction problems (Rivera et. al, 2019). Interferences such as silicon gel formation, secondary CaSO4 precipitation and secondary aluminium sulphate precipitation have been reported and should be taken into account during the processing of the slag (Davris et, 2018b).
BR sintering with Na, Ca carbonates (soda roasting) leads to the destruction of the alumino-silicate matrixes and the recovery of alumina. In such process, BR is mixed with sodium carbonate (Na2CO3) and the mixture is heated to temperatures between 800 and 1200 °C by converting the alumina into soluble phases of sodium aluminate, followed by extraction with water or mild alkaline solutions. Many research groups have applied this method for the recovery of alumina and sodium by BR (Bruckard et al., 2010; Chun et al., 2014; Kaußen and Friedrich, 2016; Li et al., 2009; Liu et al. 2012; Zhu et al., 2012). The resulting residue can be used for iron reduction by reducing smelting, while the produced slag is enriched in REE and Ti where they can be extracted with strong inorganic acids (Borra et al., 2017). Additionally, the sequence of sintering and reductive smelting process implementation does not affect the REE extraction at the final stage. Alternatively, a combined cabothermic reduction and soda roasting can also be applied to produce a magnetic iron fraction, soluble alumina phases and a residue with upgraded REE content (Cardenia et al., 2018).
A process was investigated during the gradual extraction of Fe, Al, Ti and REE (Borra et al., 2017) where initially BR sintering takes place with Na2CO 3 at 950 ° C for 4 hours followed by water leaching at 80 ° C to recover Al (75%) with further treatment of the insoluble residue by reductive smelting to recover Fe (98%) and extraction of the produced slag to recover Sc (80%) while recovery of the remaining REE is <5%. The produced slag is converted into amorphous oxides by fast cooling, while more than 80% of the Ti content, REE is recovered by leaching with 1M HCl at 25 ° C and 2% w / v pulp density. However, the presence of silica gel during the leaching process results in problematic liquid-solid separation. According to the BR sintering process with Na2CO3 and lignite (Li et al., 2014) the reduction of the hematite content to magnetite is achieved and is removed after breakage and magnetic separation. The non-magnetic fraction is extracted in 2 steps (Deng et al., 2017), firstly with H3PO4 to recover SiO2and secondly with NaOH to recover Al, while the final insoluble solid is enriched in Sc and Ti obtaining the fourfold concentration compared to the original BR.
If CaO is added to the BR reductive smelting an alkaline slag can be formed suitable for leaching alumina in alkaline environments (variation of the Pedersen process). The residue of the alkaline leaching, can then be treated with acid leaching to recover REE as well as Ti (Alkan et al., 2017, Vafeias et al. 2018)
An indicative mapping of various processing options aiming at extracting Fe, Al, and REE is presented in Figure 4.