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Selective Process Steps for the Recovery of Scandium from Jamaican Bauxite Residue (Red Mud) Remya P.N. Narayanan, Nikolaos K. Kazantzis, and Marion Heidi Emmert ACS Sustainable Chem. Eng., Just Accepted Manuscript • DOI: 10.1021/ acssuschemeng.7b03968 • Publication Date (Web): 21 Nov 2017 Downloaded from http://pubs.acs.org on December 1, 2017

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ACS Sustainable Chemistry & Engineering

Selective Process Steps for the Recovery of Scandium from Jamaican Bauxite Residue (Red Mud) Remya P. Narayanan,‡ Nikolaos K. Kazantzis,† Marion H. Emmert*‡ ‡

Department of Chemistry and Biochemistry & Center for Resource Recovery and Recycling,

Worcester Polytechnic Institute, 100 Institute Road, Worcester, MA, 01609, USA †

Department of Chemical Engineering & Center for Resource Recovery and Recycling,

Worcester Polytechnic Institute, 100 Institute Road, Worcester, MA, 01609, USA

ACS Paragon Plus Environment

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ABSTRACT. We report the development of a process allowing the selective, sustainable recovery of Scandium (Sc) with 75% efficiency from Jamaican bauxite residue (red mud), a waste product from aluminum production. The process design is inspired by green chemistry principles and focuses on establishing highly selective process steps (sulfation, leaching, and precipitation) in order to minimize costs and waste produced. In addition to Scandium oxide, the chosen approach produces mixed rare earth oxides as a side product, thus isolating an average of 88% of all rare earth elements contained in red mud. Furthermore, in light of predicted supply shortages of the critical material Sc and the need to establish cost-effective bauxite residue remediation techniques, a systematic Monte Carlo-based economic performance assessment framework is developed in order to evaluate the economic prospects of the proposed process system in the presence of irreducible uncertainty.


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Introduction Since the Chinese government introduced export restrictions for rare earths (REs; lanthanides + Sc, Y, and La) in 2011, significant price fluctuations for these critical materials have occured.1-3 However, even though prices have significantly relaxed since 2011, uncertainties in the supply of REs remain that are likely to influence current and future decisions regarding the use of RE containing materials in many applications (e.g. wind and tidal turbines, hybrid and electric vehicles). 4 Despite these uncertainties in the supply chain, the Scandium (Sc) market in particular is predicted to grow substantially over the next decades due to the projected use of Sc in light-weight, high-strength aluminum alloys in aerospace and fuel-efficient vehicle applications.5 For example, a recent market analysis predicts a significant growth of Sc consumption from a current value of 10-15 t/a in 20156 to 98 t/a for aircrafts and 3000 t/a for light vehicle production by 2023.7 As currently available Sc is mainly produced as a byproduct of other processes, increasing production rates is likely difficult;6 therefore, an urgent need for new Sc sources exists in order to prevent future shortage and price increases. Bauxite residue, also called red mud (RM), is customarily considered a waste product resulting from aluminum ore refining, but may also be viewed as a potential alternative source of Sc.8 RM is being produced and stockpiled on scales as large as 150 million t/a, of which only less than 4 million t/a find us in follow-up processes.9,10 The Sc content of RM varies between 15 and 170 ppm, depending on the source of bauxite; furthermore, other REs are also present (~850 to 1200 ppm of mostly Y, La, Ce, Pr, and Nd).11 If all Sc from RM could be recovered, 6,600 to 20,400 t/a Sc could be made available just based on the yearly production of RM. These amounts could easily address projected supply shortages of Sc, while at the same time contribute to remediating RM,


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which is dangerous to aqueous ecosystems due to its strongly basic pH.12,13 As such, using RM as a resource in a commercially viable process may also be suitable to enhance the economic performance of alumina refineries.14 Several previous attempts have been made to establish red mud as a resource; however, only few processes that establish a complete pathway from RM to Sc2O3 are known.12,15-21However, many of these processes are unlikely to be realized on a large scale due to reagent costs. Among the pieces of research work found in the relevant literature that describe various optimization approaches of single process steps, leaching studies to liberate REs from red mud are the most common.22 Unfortunately, the majority of the published studies generally only focus on recovery efficiencies and selectivities without taking reagent costs into account upon reaction design.23 Furthermore, several companies have claimed in recent years to have successfully realized the recovery of Sc from RM,24,25 but very little is known about the details of the employed process steps. With this in mind, our work focused on optimizing separation selectivity and minimizing reagent use in each step of a newly design process. Both of these principles have previously been formulated within the broader context of Green Chemistry Principles.26 Furthermore, the same design parameters have been applied for efficient and cost-effective process design in RE recovery from magnets in motors.27 Employing these principles, we identify and examine in this manuscript salient economic performance profile characteristics of the proposed process system for Sc recovery from RM (in particular 75% Sc recovery as Sc2O3). Moreover, a commercial viability analysis of the devised flowsheet shows highly favorable prospects for process commercialization, even at currently low Sc prices. Additionally, a mixed RE oxide is obtained as side product of the process, which may generate further value-enhancing opportunities. All of these exciting results


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are due to rigorous optimization efforts, which establish highly selective process steps and minimize undesired reagent and metal oxide waste.

Experimental PartAnalysis of RM. The chemical analysis of a Jamaican red mud sample was performed after dissolving the bauxite residue by an alkali fusion/acid digestion sequence: The alkali fusion was carried out by mixing 0.5 g of bauxite residue with 1.5 g of sodium carbonate and 1.5 g of sodium tetraborate decahydrate, followed by heating the mixture in an alumina crucible at 1100 °C for 30 min. The residue (3.381 g) was extracted using 100 mL 1:1 HCl:H2O and the resulting suspension was filtered. The ICP-OES analysis was performed six times to provide statistically meaningful average values of elemental contents. Sulfation & Roasting. The sulfation (acid mixing) of red mud was performed by moistening 0.50 g of dried red mud with 40 wt % water (0.20 mL) and 80 wt % concentrated H2SO4 (0.40 g, 0.22 mL) in an alumina crucible. The mixture was then heated to 120 °C for 14 h in a vacuum oven. The XRD analysis (see SI) of the obtained material confirms the completion of sulfation. The sulphated RM was then roasted at 700 °C for 1 h in a tube furnace to decompose low thermal stability sulphates to oxides. Leaching. The roasted RM (50 g) was leached by adding 50 wt-% water (25 ml) to the obtained residue, followed by ball milling (3000 rpm, glass balls; ULTRA-TURRAX® Tube Drive) for 30 min. The composition of the leach liquor was then analyzed using ICP-OES (Perkin Elmer Optima 8000) and the solid residue was analyzed by TXRF (Bruker S2 Picofox). Both analyses confirm the leaching of 84% Sc, 0% Fe, 8.2% Al, 0% Ti, 35% Ca, 100 % Y, 100 % La, 98 % Ce, 100% Pr, 99% Nd, 100% Gd, 100% Dy,68% Er, and 100 % Yb. Mixed Rare Earth Oxide Precipitation. The pH of 25 mL of the obtained leach liquor was


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adjusted to pH 8 using 2 mL of 2M NaOH. The mixed rare earth precipitate formed through pH adjustment was removed by filtration and the filtrate composition was analyzed using ICP-OES. The obtained precipitate was analyzed by TXRF. Both analyses confirm that all rare earths but Sc (Y, La, Ce, Pr, Nd, Gd, Dy, Er and Yb) are separated as a precipitate and that all Sc remains in solution in the filtrate. Scandium Oxalate Precipitation. The pH of 500 mL filtrate was adjusted to pH 1 using 15 mL of concentrated H2SO4. Then, scandium oxalate was precipitated by adding solid oxalic acid (1.0 mg; 2.02 equiv. compared to the amount of Sc in the solution). The precipitate was filtered to obtain 3.5 mg scandium oxalate from overall 10 g of RM originally leached. The precipitate was analyzed by TXRF. The filtrate was analyzed for remaining Sc by ICP-OES. Both TXRF and ICPOES analysis confirm the presence of only Sc in the precipitate and only 90% of other REs after 7 days, while leaching only small amounts of Fe and Ti (~1% and ~2%, respectively). We considered these results to be a good starting point for further studies, if (i) the recovery of Sc can be increased; (ii) leaching of bulk elements can be further decreased (even 1% Fe and 2% Ti is considered as a significant amount in the leach liquor, as both elements occur in high concentrations in RM); and (iii) the leaching times can be significantly decreased. If successful, the first process steps would be expected to result in a large volume of leach liquor; as such, RE precipitation as a second step could be expected to concentrate the desirable RE components and also serve to separate REs contained in the leach liquor from bulk elements. Finally, separation of Sc from the RE concentrate may be possible by either selective precipitation28 or solvent extraction.29 The success of this process design strategy depends strongly on the selectivities that can be achieved in each step; therefore, detailed optimization efforts for each step are described in the following sections.


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Determining RE Contents in Different Red Mud Resources. Generally, the source of RM will determine its Sc content, which in turn will likely influence the exact outcome of process optimization. Therefore, we started the process development by comparing the RE contents of different RM sources. To this end, several RM samples of different origins were collected (named in this manuscript Alcan, Alcoa, Korean, and Jamaican RM in a reflection of their origins; however, no further information regarding the source refineries or holding ponds could be obtained due to trade secrets) and analyzed. Comparison of the SEM pictures obtained for each sample showed very fine materials that form agglomerates (Figure 2).

Figure 2. SEM Images of RM Samples of Different Origins. Sources: Alcoa (top left), Jamaican RM (top right), Alcan RM (bottom left), Korean RM (bottom right).


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In a next step, bulk element concentrations were obtained for all samples, using EDX and/or ICP-OES after alkali fusion (see SI for details). Table 1 summarizes the results of this analysis. Interestingly, Al (0.05 to 0.23 kg/kg) and Ti (0.02 to 0.2 kg/kg) contents vary significantly between RM obtained from Alcoa and the other RM samples, while Fe (0.18 to 0.3 kg/kg) and Ca (0.02 to 0.03 kg/kg) contents are more comparable between samples of different origins.

Table 1. Bulk Element Concentrations for Four Types of RM as obtained by EDX and/or ICP-OES after Alkali Fusion. Element (kg/kg)

Alcoa

Jamaican RM Alcan RM

Korean RM

Fe

0.2 + 0.01

0.36 + 0.02

0.3 + 0.002

0.18 + 0.005

Ti

0.18 + 0.02

0.02 + 0.002

0.04 + 0.001

0.02 + 0.004

Ca

0.02 + 0.005

0.03 + 0.001

0.02 + 0.001

0.02 + 0.002

Na

0.059 + 0.001

0.025 + 0.001

0.14 + 0.001

0.06 + 0.001

Al

0.16 + 0.03

0.05 + 0.001

0.2 + 0.001

0.23 + 0.003

Zr

0.065 + 0.005

0.065 + 0.005

not detected

0.01 + 0.001

In order to determine the concentrations of trace elements in the different RM samples obtained, RM samples were subjected to alkali fusion and leaching: RM was first reacted with Na2CO3 and Na2B4O7·10H2O at 1100 °C, which was followed by acid digestion in HCl solution, followed by filtration (see SI for details).30 The RE concentrations obtained from leaching are depicted in Figure 3.


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400 Alcoa Jamaican

Concentration (ppm)

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300

Korean Alcan

200

100

0 Sc

Y

La Ce Pr Nd Sm Eu Gd Tb Dy Ho Er Tm Yb Lu Elements

Figure 3. Comparison of RE Element Contents of Different Red Mud Samples. The graph shows average values and standard deviations obtained from alkali fusion and acid digestion processing of at least 7 independently prepared samples.

Interestingly, the RE content varies dramatically between the samples of different origins, with Jamaican RM being the richest in both Sc (55 ppm) and other RE elements (e.g. 373 ppm Y; 287 ppm La; 366 ppm Ce; 74 ppm Pr; 70 ppm Nd; 37 ppm Gd). This suggests that Jamaican RM is the most suitable RM source obtained for Sc recovery. According to other reports, two types of RM (Australian red mud and Greek red mud) with similar or higher Sc contents (54 ppm and 120 ppm) may also be suitable as a Sc resource;21,31,32 however, these types of RM samples could not be obtained within the time frame of this study.

Selectivity in Sulfation and Roasting. Of particular importance in the recovery of Sc from RM is the selectivity in the first step, as co-leaching of REs and bulk elements can cause problems later in the recovery process, e.g. in potential solvent extraction and precipitation steps. Due to the much


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higher Fe and Ti content in comparison to the RE content in RM, entirely preventing those bulk elements from leaching is a significant challenge; at the beginning of our experimental studies, no reports that had met this challenge were available in the literature. As mentioned previously in the process design section, one report by Borra et al.19 showed results close to completely selective leaching, with only ~1% of Fe and ~2% of Ti leached from Greek RM. However, the process only recovers ~60% of Sc, while a large amount of other bulk elements (30% Ca, 100% Na) are also leached out of RM. Thus, our work focused on minimizing the amount of leached bulk elements, while maximizing Sc leaching. In an initial series of optimization steps, we focused on optimizing the sulfation/roasting sequence, while keeping the leaching conditions unchanged (H2O, 7 days, room temperature). 19 The sulfation/roasting sequence relies on the chemical reaction that converts oxides into hydrated sulfates at relatively low temperature (120 °C), which is then followed by dehydration and SO3 extrusion at higher temperatures to form oxides from unstable sulfates. As the RE-based sulfates exhibit higher decomposition temperatures than Fe sulfates (see Table 2), the roasting step would ideally transform all Fe contained in RM into insoluble oxides,23 while all REs remain soluble sulfates33-35 that can be dissolved.


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Table 2. Literature Values of Thermal Decomposition Temperatures of Relevant Metal Sulfates. Metal Sulfate

Onset Temperature for Sulfate to Oxide Conversion

Fe2(SO4)3

545 °C36

TiOSO4

340 °C37

CaSO4

>1000 °C37

Na2SO4

870 °C37

Al2(SO4)3

524 °C36

Sc2(SO4)3

700 °C38

Y2(SO4)3

850 °C39

La2(SO4)3

840 °C39

Ce2(SO4)3

666 °C36

Nd2(SO4)3

800°C39

We hypothesized that the removal of SO3 from the mixture would have a significant effect on the conversion due to chemical equilibrium considerations. Therefore, we performed all roasting optimizations in a tube furnace open to air; in contrast, the original report19 employs a muffle furnace. With this setup, the effects of different H2SO4 loadings and roasting temperatures were explored, using Jamaican RM as a starting material. Initial sulfation conditions (120 °C, 14 h) were chosen similar to the original report by Borra et al.19 Figure 4 shows the effect of roasting temperature on percent element recovery. Interestingly, the average RE recovery declines steeply at temperatures higher than 715 °C, while Fe recovery is very low when roasting is performed at temperatures higher than 700 °C.


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Figure 4. Effect of Roasting Temperature on Metal Ion Recovery in Leaching. Conditions: sulfation: 100 weight-% Jamaican RM, 40 weight-% H2O, 80 weight-% conc. H2SO4, aluminum crucible, 120 °C, 14 h, tube furnace;19 roasting: temperature specified in graphs, 1 h; leaching: H2O (solid:liquid ratio = 1/50), room temperature, 7 days.

This suggests that either 700 or 715 °C may be the most suitable temperature for selective RE recovery. Importantly, Sc recovery remains quantitative up to 715 °C (Figure 4, right side), but falls to less than 20% when roasting is performed at 750 °C. This is likely due to the decomposition of Sc2(SO4)3 under these roasting conditions; however, the demonstrated quantitative recovery of Sc up to a roasting temperature of 715 °C is a significant advancement in comparison to the reported conditions. We speculate that this is possible due to the improved furnace setup used for the herein described studies, which removes gaseous products from the chemical equilibrium, thus allowing a higher-yielding formation of insoluble Fe oxides and higher Sc selectivity. In a next step, the effect of the used H2SO4 loading on Sc recovery and selectivity was studied. We chose to perform this study at a roasting temperature of 700 °C to maximize the overall RE recovery (88%) at a relatively low Fe recovery level (5%; see Figure 4 above, obtained with 80 weight-% of conc. H2SO4 used for sulfation and roasting). Interestingly, employing higher H2SO4


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loadings (100%), provided the same 88% overall RE recovery, but Fe recovery was further enhanced (34%), thus diminishing the selectivity for RE recovery (Figure 5). This may be due to the incomplete decomposition of Fe2(SO4)3 or to a more complete conversion of Fe oxides to Fe2(SO4)3 during the sulfation reaction. Due to the lower obtained selectivity, the use of higher H2SO4 loadings than 80 weight-% was considered unfavorable for further development.

Figure 5. Influence of H2SO4 Loading on Leaching Selectivity and Yield. Conditions: sulfation: 100 weight-% Jamaican RM, 40 weight-% H2O, 60-100 weight-% conc. H2SO4, aluminum crucible, 120 °C, 14 h, tube furnace; roasting: 700 °C, 1 h; leaching: H2O (solid:liquid ratio = 1/50), room temperature, 7 days.

Figure 5 further shows that lower H2SO4 loadings significantly decreased RE recovery (only 36% average) and Sc recovery in particular (0%). Despite the low Fe leaching observed under these conditions (0%), they were thus deemed unfavorable. In summary, sulfation with 80 weight% conc. H2SO4 followed by roasting at 700 °C for 1 h were the optimized sulfation/roasting


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conditions, providing the maximum average RE recovery (88%, with 100% Sc recovery) along with recovery of 4.9% Fe, 3.0% Ti, 22.4% Ca, 21.7% Al, and 70% Na.

Selectivity and Efficiency in Leaching. The above detailed optimization steps had achieved one important goal: Leaching Sc and other REs nearly quantitatively from Jamaican RM. To further improve the selectivity of RE leaching vs. bulk element leaching, the next series of studies focused on optimizing the leaching step itself. To this end, initial experiments were performed in order to gain insight into the leaching kinetics to determine if selectivity can be achieved by adjusting leaching times. As is illustrated in Figure 6, leaching of the solid residue obtained from roastin at room temperature is a slow process: The maximum RE recovery is only obtained after 6 days. Furthermore, the selectivity of RE leaching as compared to bulk element leaching does not considerably change during the course of 7 days. However, processing times of this magnitude are difficult to realize in pilot or production plants; as such, we turned our efforts to approaches that may possibly enable faster leaching.


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Figure 6. Kinetics of Metal Ion Leaching. Conditions: sulfation: 100 weight-% Jamaican RM, 40 weight-% H2O, 80 weight-% conc. H2SO4, aluminum crucible, 120 °C, 14 h, tube furnace; roasting: 700 °C, 1 h; leaching: H2O (solid:liquid ratio = 1/50), room temperature, 1 to7 days.

We rationalized that higher leaching temperatures may accelerate leaching rates by providing more thermal energy to overcome activation barriers required for hydrating RE ions contained in a solid matrix. However, leaching experiments at 50 or 80 °C led to even slower leaching of REs from the solid residue, compared to the leaching behavior at room temperature (Figure 7). This may be due to the lower solubility of RE sulfates at higher temperatures.20 As such, leaching at higher temperatures was considered unsuitable for improving the process.


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Figure 7. Kinetics of RE Leaching at Different Leaching Temperatures. Conditions: sulfation: 100 weight-% Jamaican RM, 40 weight-% H2O, 80 weight-% conc. H2SO4, aluminum crucible, 120 °C, 14 h, tube furnace; roasting: 700 °C, 1 h; leaching: H2O (solid:liquid ratio = 1/50), room temperature/50 °C/80 °C, 1 to 6 days.

The formation of small, co-crystalline morphologies in RM after acid digestion has previously been documented;40 therefore, we speculated that microcrystallinity of the residue obtained by roasting may hinder liberation of REs from the crystal matrix, thus leading to long leaching times. We further reasoned that leaching from co-crystalline materials may be accelerated by applying force to the material, which could be applied by sonication or milling. Therefore, our next optimization study measured leaching performance upon sonicating (Figure 8, left).


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Figure 8. Leaching Upon Sonication of Leaching Mixture. Conditions: sulfation: 100 weight% Jamaican RM, 40 weight-% H2O, 80 weight-% conc. H2SO4, aluminum crucible, 120 °C, 14 h, tube furnace; roasting: 700 °C, 1 h; leaching: H2O (solid:liquid ratio = 1/50), room temperature, sonication in water bath/no sonication.

Excitingly, the average RE leaching observed was 80% after only 1 h of sonication, with almost quantitative RE leaching after 5 h. Moreover, the leaching of bulk elements was significantly decreased upon sonication leaching (to 0% Fe, 6% Ti, 22% Ca, and 6% Al). This is an exciting discovery, as Fe and Ti are present in large quantities in RM and may interfere in RE isolation from the leach liquor. Sonication has previously been documented to lead to a reduction in particle size of crystalline material.41 We speculate that a reduction of particle size of the roasted RM residue through sonication may accordingly result in larger surface areas at which ion exchange with the surrounding leach liquor can occur, thus accelerating the leaching process. . The right part of Figure 8 contrasts RE leaching kinetics of the sonication method with the kinetics of traditional, stirred RE leaching – a clear improvement can be observed. Similar to sonication, ball milling also resulted in a significant increase in leaching rate and selectivity (Figure 9): after only 30 min, 88% average RE recovery (89% Sc) was obtained in the absence of any Fe and Ti leaching and at relatively low Al and Na leaching (10% and 10%,


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respectively). Extending the ball milling time to 60 min increased the recovery of REs to 95%; however, the recovery for Sc remained at 89%, while Fe and Ti leaching (0.01 and 1.06%, respectively) started to occur. The exact mechanisms for achieving such highly selective Sc leaching are currently not understood in detail. In-depth understanding will likely require further detailed studies, which were considered outside the scope of this manuscript.

Figure 9. Metal Ion Leaching after 30 min (left) and 60 min (right) upon Ball Milling. Conditions: sulfation: 100 weight-% Jamaican RM, 40 weight-% H2O, 80 weight-% conc. H2SO4, aluminum crucible, 120 °C, 14 h, tube furnace; roasting: 700 °C, 1 h; leaching: H2O (solid:liquid ratio = 1/50), room temperature, ball milling (IKA ULTRA-TURRAX® Tube Drive, glass balls diameter 6 mm, 3000 rpm), 30 or 60 min.

Finally, the influence of the solid/liquid ratio on the selectivity and efficiency of leaching was investigated; we reasoned that a lower volume of leach liquor would be desirable for further processing. However, lowering the amount of H2O used for leaching (Figure 10) resulted in a significant decrease of overall RE recovery from 95% (with 50 weight-% H2O) to 69%, 65%, and


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61% (with 30, 20, or 10 weight-% H2O). We speculate that this decrease in leaching efficiency may be caused by solubility limitations of the leached salts in the leach liquor. This decrease in RE recovery was undesirable; therefore, further experiments were performed with a leach liquor obtained after 30 min of ball milling with a solid/liquid ratio of 1:50 (w/w). As a result of these studies, the leach liquor contains the relevant metal concentrations as shown in .

Figure 10. Ball-Mill Supported Leaching of RM Residue at Different Solid/Liquid Ratios. Conditions: sulfation: 100 weight-% Jamaican RM, 40 weight-% H2O, 80 weight-% conc. H2SO4, aluminum crucible, 120 °C, 14 h, tube furnace; roasting: 700 °C, 1 h; leaching: H2O (1.0 weight% solid residue, 10-50 weight-% H2O), room temperature, ball milling (IKA ULTRA-TURRAX® Tube Drive, glass balls, 3000 rpm), 60 min. Table 3. Element Concentrations of Leach Liquor.a Element

Concentration (g/L)

Na

3.71 + 0.0005

Al

2.60 + 0.005

Ca

6.06 + 0.005

Sc

0.055 + 0.008


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Y

0.363 +0.001

La

0.317 +0.001

Ce

0.406 +0.002

Pr

0.082 +0.001

Nd

0.078 +0.001

Sm

N.D

Eu

N.D

Gd

0.032+0.001

Tb

N.D

Dy

0.034+0.0001

Ho

0.009+ 0.0007

Er

0.011+0.0006

Tm

0.01+0.0001

Yb

0.012 +0.0005

a

Conditions: sulfation: 100 weight-% Jamaican RM, 40 weight-% H2O, 80 weight% conc. H2SO4, aluminum crucible, 120 °C, 14 h, tube furnace; roasting: 700 °C, 1 h; leaching: H2O (1.0 weight-% solid residue, 10-50 weight-% H2O), room temperature, ball milling (IKA ULTRA-TURRAX® Tube Drive, glass balls, 3000 rpm), 60 min.

Selectivity for RE Precipitation. The obtained RE solution (leach liquor) contained all REs, including Sc and Y, as well as La, Ce, Pr, Nd, Gd, Dy, Er, and Yb. Typical literature approaches to separate and concentrate the REs have used solvent extraction methods,42 which typically use relatively expensive organic extractants and solvents. Therefore, a direct use of solvent extraction with the obtained, dilute leach liquor as aqueous phase was deemed undesirable from a cost perspective. We postulated that precipitation of the REs contained in the solution may allow


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concentrating the valuable components while decreasing the volume of the solutions for further processing (Figure 11).

Figure 11. Weight Balance and Further Separation Options after Roasting and Leaching.

Thus, we explored conditions for precipitation by adjusting the pH of the leach liquor with NaOH. The original pH of the obtained solution was measured to be 3.57; addition of NaOH until a pH of 8, 10, 12, and 14 was measured provided a precipitate in each case. Determining the remaining RE concentrations in solution by ICP-OES determined that the majority of REs had precipitated quantitatively even at pH 8 (Figure 12), with some Ca also co-precipitating. However, Sc does not precipitate under these conditions, allowing for separation of Sc from the other REs at this stage of the process. This separation can be rationalized with the amphoteric behavior of Sc(OH)3, which shows significant solubility at both low and high pH values;43 similar amphoteric behavior is known for Al(OH)3,44 which also reports to the filtrate. Furthermore, analysis of the precipitate by TXRF confirmed that only traces of Sc were contained in the formed solid. These data suggest that precipitation does not only concentrate REs, but that this step also allows for an efficient separation of Sc from the other REs originally contained in RM. Interestingly, no


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precipitation of Al occurs under these conditions, suggesting that Al is separated together with Sc from the REs through precipitation.

Figure 12. RE Precipitation from Leach Liquor. Leach liquor obtained after filtering off bulk materials after sulfation, roasting, leaching (30 min, ball milling); pH of leach liquor adjusted with NaOH (2M). Precipitation percent are obtained by analyzing the filtrate by ICP-OES.

Considering the mass balance (Figure 13, Table 3), it is inferred that the NaOH precipitation step provides access to a valuable, mixed RE oxide side product (mixed with Ca oxide). Furthermore, Sc is completely separated in the precipitation step (see concentrations in leach liquor, Table 3). This simplifies the separation of Sc in the next step, as only the bulk elements Na, Ca, and Al have to be considered. Similarly to the first highly selective leaching step, high selectivity in precipitation thus improves the overall efficiency of the process.

Table 4. Element Concentrations of Filtrate after Precipitation.a Element

Concentration (g/L)


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Sc

0.051 +0.001

Ca

2.045 +0.002

Al

2.606 +0.001

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a

Conditions: sulfation: 100 weight-% Jamaican RM, 40 weight-% H2O, 80 weight% conc. H2SO4, aluminum crucible, 120 °C, 14 h, tube furnace; roasting: 700 °C, 1 h; leaching: H2O (1.0 weight-% solid residue, 10-50 weight-% H2O), room temperature, ball milling (IKA ULTRA-TURRAX® Tube Drive, glass balls, 3000 rpm), 60 min.

Figure 13. Mass Balance and Separation Obtained By Precipitation.

Selectivity in Sc Precipitation. Due to the Sc separation obtained in the precipitation step at pH 8, the remaining solution contains the vast majority of Sc leached from RM in addition to Na, Ca, and Al. In agreement with the literature,45,46 precipitating Sc from this solution by NaOH addition to directly obtain Sc(OH)3 was unsuccessful, suggesting that other precipitation reagents need to be explored. Previous studies suggest that Sc can be precipitated from aqueous solutions as the oxalate salt upon addition of oxalic acid.24,47 We considered oxalic acid a desirable reagent,


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because oxalate salts can be converted into oxides simply by roasting.48 Furthermore, oxalic acid is a solid reagent, and therefore less hazardous than other acidic reagents. When oxalic acid precipitation was attempted with the solution obtained after filtering off the mixed RE/Ca precipitate (pH 8), several interesting observations were made (Figure 14, top left chart): (i) Ca co-precipitates with Sc under these conditions; (ii) the Sc/Ca selectivity increases with the amount of oxalic acid equivalents added; (iii) the recovery of Sc also increases when more oxalic acid is added. The latter two observations both suggest that Sc precipitation selectivity may improve under more acidic conditions.

Figure 14. Optimization of Oxalic Acid Precipitation of Sc. Conditions: pH

Selective Process Steps for the Recovery of ... - ACS Publications

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