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.