A Novel Dissolution and Synchronous Extraction of Rare Earth Elements from Bastnaesite by a Functionalized Ionic Liquid [Hbet][Tf₂N]

Abstract

Rare earth elements (REEs) are widely used in high-tech industries and are important basic raw materials. Bastnaesite is one of the most important minerals used in the rare earth extraction industry, and the efficient development of it is the key guarantee for the safe supply of rare earth raw materials. In this study, a novel method for dissolving bastnaesite using a carboxyl-functionalized ionic liquid ([Hbet][Tf₂N]) is proposed. This innovative method provides a collaborative model with the dissolution and synchronous extraction of rare earth elements during the heating and cooling of the [Hbet][Tf₂N]–H₂O system. In the heating process, rare earth elements can be dissolved in a weakly acidic environment of ionic liquid without the trouble of HF escaping, and the leaching efficiencies of rare earth elements are above 95%. During the cooling of the leaching system, the rare earth ions in the dissolved state are extracted into the ionic liquid phase due to the two-phase stratification of [Hbet][Tf₂N] and aqueous solution. It has been proved that rare earth ions recovery and ionic liquid regeneration can be achieved by back extraction using oxalic acid for the REEs-loaded ionic liquid.

1. Introduction

Rare earth elements (REEs) play an irreplaceable role in high-tech applications for their special electron layer structure and unique physical and chemical properties, such as superconductivity, fluorescence, laser, permanent magnetism, polishing, catalysis, and hydrogen storage [1,2,3,4,5,6,7]. Now, there are still no substitutes for the fifteen REEs such as lanthanum, yttrium, and scandium, and they are gradually becoming important support materials for the high-technology sector. The global demand for REEs is still growing by 5% a year.

Bastnaesite is the most abundant light rare earth mineral in the world, in which REEs mainly exist in the form of rare earth fluorocarbonates. After the beneficiation process, the bastnaesite concentrate can generally be obtained with a grade of 50% to 70% (the content of REO). Currently, the main treatment processes for bastnaesite include oxidation roasting–acid leaching and sodium hydroxide decomposition–hydrochloric acid leaching [8,9,10,11,12,13]. Hydrometallurgical processes for bastnaesite are still plagued by the element fluorine [14,15,16], especially in the acid-leaching step [17,18,19]. When using inorganic acid to leach bastnaesite or its decomposition products, high leaching acidity will easily lead to the escape of hydrogen fluoride, and low leaching acidity will lead to a low recovery of rare earth. The effect of fluorine on the recovery of rare earth is inevitable in inorganic acid leaching because of the strong coordination between fluorine and rare earth.

Compared with the inorganic acid in the traditional hydrometallurgical leaching process, the functionalized acidic ionic liquids, molten salt at room temperature, are potential substitutes for hydrometallurgical leaching, which has the advantages of low acidity of the leaching system and excellent metal ion complexation ability. They are also considered green reagents due to their low volatility [20,21,22,23,24,25,26]. Moreover, the ionic liquid composed of cations and anions can be designed artificially according to the functional needs of the types and the functional groups of the cations and anions [27], which can give ionic liquids with special functions, such as acidity and alkalinity, hydrophobicity and hydrophilicity [28,29,30,31]. However, the application of ionic liquids in hydrometallurgy still accounts for a small proportion, and its great potential needs to be further explored.

Betainium bis(trifluoromethylsulfonyl)imide ([Hbet][Tf₂N]) was synthesized by Nockemann et al. in 2006 [32]. Because the cation of [Hbet][Tf₂N] contains a carboxyl functional group, most of the metal oxides can be dissolved in the system of [Hbet][Tf₂N]-H₂O but not iron, silicon, and aluminum oxides [32,33,34]. In addition, [Hbet][Tf₂N] has unique theriomorphic behavior with water [32,35,36], and the mutual dissolution of ionic liquids and water depends on temperature. Above the critical mutual solubility temperature, [Hbet][Tf₂N] and water dissolve into one phase, and [Hbet][Tf₂N] and water will separate into two phases with clear interfaces, just like the organic extractant and aqueous solution in the extraction process. This feature gives it the advantage of dissolving metal compounds and extracting metal ions simultaneously during the heating and cooling of the [Hbet][Tf₂N]-H₂O system [37,38,39,40]. Thus, [Hbet][Tf₂N] has been used to recycle rare earth from waste lamp phosphor powders [41], and the recovery of key rare earth elements could reach 80%. [Hbet][Tf₂N] was also used to purify indium hydroxide through a dissolution–extraction process [42], and the separation effect of indium and impurity metal ions (Al³⁺, Ca²⁺, Cd²⁺, Ni²⁺, Zn²⁺) is remarkable. In addition, [Hbet][Tf₂N] loaded with metal ions can also be regenerated by an acidic solution, just like extractants [37,43,44,45]. Based on the characteristics of [Hbet][Tf₂N], the short process of dissolution and simultaneous extraction will have great development prospects.

In this study, the ionic liquid [Hbet][Tf₂N] is applied for leaching bastnaesite, one of the most important light rare earth minerals. The effect of fluoride on the leaching efficiency of rare earth was avoided by the leaching–extraction of REEs by [Hbet][Tf₂N]. REEs ions were enriched into ionic liquids by exploiting the temperature-controlled properties of the ionic liquid. Finally, the oxalic acid solution was used to strip the [Hbet][Tf₂N] loaded with REEs, in which the recovery of REEs and the regeneration of [Hbet][Tf₂N] could be realized.

2. Materials and Methods

2.1. Chemicals and Materials

Betaine hydrochloride (HbetCl) (99.5%) and Lithium bis(trifluoromethylsulfonyl) imide (LiTf₂N) (99.0%) were purchased with an AR degree from Alfa Aesar (Zhengzhou, China). Sodium hydroxide (NaOH) (96%) with an AR degree was purchased from Yongda Chemical Reagent Company Limited (Tianjin, China). Oxalic acid dihydrate (H₂C₂O₄·2H₂O) (99.8%) with a GR degree; silver nitrate (AgNO₃) (99.8%) with an AR degree; and standard chloride solutions of cerium, praseodymium, neodymium, and lanthanum (1 g/L) were purchased from Sinopharm Chemical Reagent Co., Ltd. (Shanghai, China). Hydrochloric acid (HCl) solutions were prepared using concentrated HCl (37%), and all solutions were prepared using ultrapure water obtained from a Best-R ultrapure water system. The above chemical reagents were used as received without further purification.

2.2. Instrumentation and Methods

Metal ion concentrations were determined using an Inductively Coupled Plasma Optical Emission Spectrometer (ICP-OES, PerkinElmer, Waltham, MA, USA). X-ray diffraction (XRD) patterns of the samples were obtained using a Philips PW3040/60 powder diffractometer with Cu Kα radiation. Microstructures were obtained using a scanning electron microscope (SEM, FEI QUANTA 200, Eindhoven, The Netherlands) with an energy-dispersive X-ray spectroscopy (EDS) detector. Fourier transform infrared spectroscopy (FT-IR) spectra were recorded using a Thermo Scientific Nicolet iS5 spectrometer over a frequency range of 400–4000 cm⁻¹. The Thermal Analyzer (STA 449 F5, NETZSCH, Selb, Bavaria, Germany) was over a temperature range of 30 °C–1000 °C. Raman spectra were recorded using a LabRAM HR Evolution (Raman, Horiba, Japan). The viscosity of the ionic liquid was measured with a Rheometer (Anton Paar MCR 302, Graz, Austria).

2.3. Synthesis Method of Ionic Liquid

The ionic liquid betaine bis(trifluoromethylsulfonyl) imide ([Hbet][Tf₂N]) was synthesized with betaine hydrochloride and lithium bis (trifluoromethylsulfonyl) imide based on the one-step method. HbetCl (0.2 mol, 30.72 g) and LiTf₂N (0.2 mol, 57.42 g) were mixed and dissolved in deionized water (50 mL), and the mixed solution was then placed at room temperature and stirred magnetically at 800 rpm for 1 h. After the reaction was completed, the solution was left to separate naturally into two phases. The aqueous phase containing LiCl was separated from the ionic liquid using a separating funnel. To make the ionic liquid free of chloride ions, the ionic liquid was washed with deionized water several times. The water was removed from ionic liquids by using a rotary evaporator. The reaction equation for the synthesis of ionic liquid is shown in Equation (1). The structural formula of the ionic liquid is shown in Figure 1.

HbetCl + LiTf₂N=[Hbet][Tf₂N] + LiCl

(1)

2.4. Experimental Process

The process flow chart of the bastnaesite treatment is shown in Figure 2, which mainly includes the alkaline decomposition pretreatment for bastnaesite, leaching, and extraction with ionic liquid, and the back extraction of [Hbet][Tf₂N] loaded with REEs.

2.4.1. Dissolution and Synchronous Extraction of Rare Earth Elements from Bastnaesite

To enhance the leaching effect of bastnaesite, the bastnaesite was pretreated by alkali decomposition, in which the rare earth fluorinated carbonate was converted into rare earth hydroxide. After the alkali decomposition treatment, the solid product washed with deionized water several times was dried at 70 °C for 3 h. A certain volume ratio of [Hbet][Tf₂N] and deionized water was placed in a conical flask. When the [Hbet][Tf₂N]–H₂O system was heated to the set temperature in a water bath, a certain mass of alkali decomposition products was added with the liquid–solid ratio in the range from 8 to 40 mL/g. After the end of the leaching experiment, the leaching was cooled to room temperature. Then, the leaching residue and the solution were separated using a centrifuge at 10,000 rpm for 10 min. The metal ion concentration in the solution was analyzed using ICP-OES. The fluoride ion content was then measured with the fluoride ion selective electrode method. The leaching efficiency of REEs and fluoride were calculated using Equation (2):

$$\eta =\frac{C\times N\times V}{m\times \omega }$$

(2)

ղ is the percentage dissolution of REEs and fluoride. C is the mass concentration of REEs and fluoride measured by ICP-OES and the fluoride ion selective electrode method. N is the dilution times of the solution before measuring. V is the total volume of the solution. m is the quality of the sample in the experiment. ω is the mass percentage of REEs and fluoride in the bastnaesite of alkali decomposition products.

The concentration of REEs in the solution was analyzed using ICP-OES, and the analysis of the fluoride ion concentration in the solution was performed using a fluoride-ion-selective electrode.

2.4.2. Back Extraction of Rare Earth Elements

The recovery of REEs and the regeneration of [Hbet][Tf₂N] were accomplished by stripping in the [Hbet][Tf₂N] loaded with REEs with the oxalic acid solution. A solution of [Hbet][Tf₂N] loaded with REEs and oxalic acid was mixed, and the mixture was placed in a water bath with magnetic stirring. After the reaction was completed, it was allowed to stand for a period of time, and then the aqueous phase and the ionic liquid phase were separated, and the solid product and the solution were separated by centrifugation at 10,000 rpm for 10 min. The concentration of the remaining rare earth elements in the solution was analyzed using ICP-OES to calculate the reverse extraction rate. After washing and drying the solid products, the physical phases as well as the microstructure were analyzed.

3. Results and Discussion

3.1. Alkali Decomposition of Bastnaesite

The chemical composition, phase, and microstructure analysis results of bastnaesite are shown in

Table 1. Chemical composition of bastnaesite.

Elements	Ce	La	Nd	Pr	Ca	P	F
Wt%	19.98	15.95	7.41	4.79	0.37	0.02	7.96

and Figure 3. Bastnaesite mainly contained cerium, lanthanum, neodymium, and praseodymium, and the main phase is rare earth fluoride carbonate. Bastnaesite particles showed a regular dense surface bulk morphology, which is very unfriendly to the leaching process using the low acid ionic liquid. Therefore, the alkali decomposition method was used to destroy the pyknotic surface of the particles to enhance the leaching effect. The alkaline decomposition process can not only promote the conversion of the rare earth phase into rare earth hydroxide but also converts fluorine into sodium fluoride for removal. In other words, the removal ratio of fluoride can also represent the decomposition effect of bastnaesite.

The effect of different sodium hydroxide concentrations and different reaction durations on the alkali decomposition of bastnaesite was investigated (Figure 4). It can be seen from Figure 4a that when the concentration of sodium hydroxide was 30%, the removal ratio of fluorine reached the maximum. Moreover, it can be seen from Figure 4c that the 20% sodium hydroxide solution could not completely decompose bastnaesite, as observed in the obvious diffraction peaks of bastnaesite. Figure 4b showed that the fluorine removal ratio increased with the increase in reaction duration, and it leveled off after four hours. In addition, the products were mainly composed of rare earth hydroxides. The SEM-EDS analysis results of the alkali decomposition product are shown in Figure 5. Compared to bastnaesite (Figure 3), the morphology of the mineral particles changed obviously, and the surfaces on the particles were loose and porous and many rod-like crystals were deposited on the surface. The transformation of the morphology and phase during alkali decomposition is beneficial to the leaching of REEs by [Hbet][Tf₂N].

3.2. Leaching Behavior of REEs from Alkali Decomposition Products

Rare earth hydroxides were more easily dissolved in the [Hbet][Tf₂N]–H₂O system than fluorinated carbonates. Based on the results of thed alkali decomposition pretreatment, the effects of the volume ratio ([Hbet][Tf₂N] to H₂O), liquid–solid ratio (L/S), reaction temperature, and time on the leaching efficiency of cerium, praseodymium, neodymium, and lanthanum were investigated by single-factor experiments, and the results are shown in Figure 6.

As seen in Figure 6a, the leaching efficiencies of REEs reached their maximum when the L/S was 40 mL/g. Due to the increase in the solid phase, the consumption of hydrogen ions increased, and the in-and-out effect was weakened. As shown in Figure 6b, the leaching efficiency of REEs in the leaching system increased with the volume of [Hbet][Tf₂N]. The leaching effect of REEs was closely influenced by the acidity and viscosity of the leaching system, as they were both related to the content of [Hbet][Tf₂N], whereby they formed a positive relationship. When the water content reached the required amount, the increase in ionic liquid was more favorable for leaching. Temperature was a significant factor affecting the viscosity of [Hbet][Tf₂N] as the viscosity of pure [Hbet][Tf₂N] was very large, approaching 1000 mPa·s at room temperature. With the increase in temperature and water content, the viscosity obviously reduced as shown in Figure 7, which is conducive to the progress of the leaching process. In general, when the temperature was higher than the critical temperature (55 °C), the leaching effect of rare earth was better. However, the solubility of rare earth hydroxide in IL-H₂O was high and easy, resulting in little effect of temperature on leaching efficiency. Figure 6d shows that the leaching efficiency of rare earth gradually increased with the increase in the reaction duration. When the leaching time reached 90 min, it gradually stabilized. Under different leaching conditions, the leaching efficiency of fluorine in the [Hbet][Tf₂N]–H₂O system was small and could basically be ignored. It can be considered that the undecomposed rare earth fluorocarbonate was not leached.

The alkali decomposition products could not be completely dissolved in the [Hbet][Tf₂N]–H₂O system. The leaching residues were further analyzed to determine the composition and morphology of the insoluble residues.

SEM-EDS results of the leaching residues showed that (Figure 8a–c) the particles of the insoluble residues showed the microstructure of corrosion, which is completely different from the bastnaesite and alkali decomposition products. The XRD and EDS results indicated that the residues were mainly undecomposed bastnaesite.

FTIR curves (Figure 8e) of bastnaesite, alkali hydrolysis products of bastnaesite, and leaching residue showed that the absorption peaks of the leaching residues and bastnaesite were basically the same at 1086 cm⁻¹, 870 cm⁻¹, 1456 cm⁻¹, and 727 cm⁻¹, which correlate to the symmetric stretching vibration peak, out-of-plane bending vibration peak, asymmetric stretching vibration peak, and in-plane bending vibration peak of CO₃²⁻, respectively. The absorption peaks at 3606 cm⁻¹ and 1644 cm⁻¹ of the bastnaesite alkaline hydrolysis products were -OH stretching vibration peaks. It had an absorption peak of CO₃²⁻ at 1456 cm⁻¹, indicating that there was still a small amount of bastnaesite in the sample that had not been completely hydrolyzed.

From the Raman spectra (Figure 8f), the peak positions of bastnaesite at 255 cm⁻¹, 346 cm⁻¹, and 393 cm⁻¹ were related to halogen fluorine. Its peak positions at 686 cm⁻¹ and 737 cm⁻¹ were the in-plane bending vibration peaks of CO₃²⁻, and the peak position at 868 cm⁻¹ was the out-of-plane bending vibration peak of CO₃²⁻. After alkali hydrolysis of bastnaesite, the peaks of -OH of rare earth hydroxides were detected at 1608 cm⁻¹ and 3598 cm⁻¹. After leaching, the hydroxide was dissolved, and the rest was mainly the remaining fluorocarbonate due to incomplete alkaline decomposition, so its Raman spectrum was similar to that of bastnaesite. This result indicated that the leaching efficiencies of REEs were obviously affected by the alkaline decomposition pretreatment of bastnaesite.

3.3. Extraction of REEs Based on Stratification of [Hbet][Tf₂N] and Aqueous Solution

The aqueous phase and the [Hbet][Tf₂N] phase separated to form two phases with well-defined interfaces when the leachate temperature was lower than the critical temperature. After the separation of the aqueous phase and IL phase of the leaching solution in room temperature, the partitioning equilibrium of rare earth elements in the two phases was studied, as shown in Figure 9. Figure 9a shows that the distribution of REEs (Ce, Pr, Nd, and La) changed with the water content. As the water content decreased, more REEs were distributed in the ionic liquid. The increase in the solid-to-liquid ratio led to an increase in the metal ions in the leching solution. Therefore, the rare earth ion distribution ratio in the ionic liquid phase increased. Similarly, the increase in the leaching duration led to an increase in the content of rare earth elements and thus to a greater distribution of metal ions in the ionic liquid. The REEs content of the leaching solution varied little at different temperatures, so the leaching temperature had a weak effect on its distribution.

3.4. Back Extraction of [Hbet][Tf₂N] Loaded with REEs

The dissolved Ce³⁺, Pr³⁺, Nd^3+, and La³⁺ in [Hbet][Tf₂N] was recovered by a back extraction with the oxalic acid solution, and the regeneration of [Hbet][Tf₂N] was realized. The acidic aqueous phase and [Hbet][Tf₂N] loaded with REEs ions were mutually dissolvable when heated above the critical temperature. During this process, rare earth ions (Ce, Pr, Nd, La) were separated as precipitates (e.g., oxalate). The acidic proton protonated the betaine ligand and regenerated the ionic liquid (Equation (3)).

2[M(bet)₃][Tf₂N]+3H₂C₂O₄=M₂(C₂O₄)₃+6[Hbet][Tf₂N]

(3)

The influence factors (acid concentration, volume ratio of acid solution and [Hbet][Tf₂N], reaction temperature, and duration) of stripping were investigated. The experimental results are shown in Figure 10.

According to the calculation, the theoretical dosage of oxalic acid should be 1.5 times the content of rare earth ions. When the concentration of oxalic acid is less than 0.8 M, the amount of oxalic acid is less than the theoretical amount, and thus the recovery of REEs is low. Enough oxalic acid can completely precipitate rare earth ions, so the optimal oxalic acid concentration is 1.0 M. When the volume ratio of the oxalic acid solution to [Hbet][Tf₂N] increased to 1, the amount of oxalic acid was greater than the theoretical amount and the recovery of REEs was close to 100%. The presence of metal ions can significantly reduce the mutual dissolution temperature. The addition of oxalic acid can reduce the metal content in the system, resulting in the influence of its mutual solubility temperature. Therefore, the two phases are not completely miscible with each other even if the temperature is below 45 °C, which results in the reaction between oxalic acid and rare earth ions being incomplete. The optimum reaction temperature is 65 °C. As shown in Figure 10d, the reaction between the [Hbet][Tf₂N] and oxalic acid was relatively rapid, and the stripping of REEs could be achieved within a short time. To sum up, the optimal conditions for back extraction were 1 M oxalic acid, a volume ratio of oxalic acid to [Hbet][Tf₂N] of 1, 65 °C, and a reaction duration of 10 min.

The white solid product and the regenerated [Hbet][Tf₂N] in the back extraction process were analyzed. The FIIR curve of regenerated [Hbet][Tf₂N] (Figure 11) was consistent with that of the original ionic liquid, and the regeneration of [Hbet][Tf₂N] could be used to leach the REE again.

The white solid products were collected and roasted at 200 °C and 600 °C for 1 h. Then, the products and roasted products were analyzed with XRD, FTIR, and SEM-EDS. As shown in Figure 12a, the main ingredient of the precipitations was REEs (Ce, Pr, Nd, La) oxalate hydrate, and it was in the form of flakes (Figure 13a). Owing to the loss of crystal water, the roasted products at 200 °C became amorphous (Figure 12a). It is also suggested in Figure 13d that the content of the oxygen element decreased. However, FTIR (Figure 12b) showed that there were the same characteristic peaks between the low temperature (200 °C) roasted products and REE oxalate hydrate, except the -OH peak of water at 3420 cm⁻¹. The XRD and FTIR curves of the roasted products at 600 °C suggest that the high temperature (600 °C) roasted products were the mixed REE oxidation. Figure 13f shows that the content of carbon was considerably reduced. In addition, there was almost no change in the morphology before and after roasting.

4. Conclusions

This study aimed to propose a novel method for dissolving bastnaesite using a carboxyl-functionalized ionic liquid ([Hbet][Tf₂N]). This method provides a collaborative model with dissolution and synchronous extraction of rare earth elements during the heating and cooling of the [Hbet][Tf₂N]–H₂O system. Under the conditions of the volume ratio of [Hbet][Tf₂N]–H₂O 8:12, L/S 8 mL/g, temperature 75 °C, and a leaching duration of 60 min, the leaching efficiency of the four rare earth elements could reach approximately 90%. During the leaching process, the bastnaesite remaining due to incomplete alkaline hydrolysis was not dissolved, and the fluorine in it did not interfere with the leaching process. By means of the delamination of [Hbet][Tf₂N] and aqueous solution after cooling, REEs in the leachate were mostly enriched in the ionic liquid phase, which is equivalent to the extraction of REEs from an aqueous solution. The use of the oxalic acid solution could strip the rare earth ions supported on the IL and regenerate the ionic liquid at the same time. REEs could be totally stripped using a 1 M oxalic acid solution. Particularly, oxalic acid stripping can be used to directly obtain rare earth oxalate precipitation. The ionic liquid can be reused for leaching after regeneration. Therefore, this approach exhibits efficient leaching and extraction, demonstrating that [Hbet][Tf₂N] is effective at recovering REEs from bastnaesite with the dissolution and synchronous extraction of REEs.