Kinetic Mechanisms and Efficient Leaching of Praseodymium, Neodymium, Fluorine, and Lithium from Molten-Salt Slag via Atmospheric Alkaline Leaching

Abstract

Rare-earth molten-salt electrolysis slag contains a substantial quantity of rare-earth elements, rendering it a valuable secondary resource for rare-earth recovery. To achieve the efficient recovery of praseodymium (Pr), neodymium (Nd), lithium (Li), and fluorine (F) from rare-earth molten-salt electrolysis slag, this paper proposes an atmospheric alkaline leaching method. The leaching efficiency of Nd, Pr, F (95.02%), and Li (95.87%) can be reached at a NaOH concentration of 80%, a reaction temperature of 180 °C, a reaction time of 2 h, and an alkali to slag ratio of 3:1. Leaching efficiency kinetic analysis shows that the leaching processes of fluorine and lithium are both controlled by interfacial chemical reactions, with apparent activation energies of 59.06 kJ/mol and 57.33 kJ/mol, respectively. The mineral phase transformation and morphological analysis were studied by X-ray diffractometer and scanning electron microscope. The results indicated that rare-earth fluoride (REF₃) reacts with sodium hydroxide to form rare-earth hydroxide (RE(OH)₃) and soluble sodium fluoride (NaF), while LiF is converted into LiOH and enters the liquid phase. High-efficiency separation was achieved by washing with water, avoiding high-temperature energy consumption and the problem of fluorine-containing waste gas.

1. Introduction

Rare earths, including, dysprosium, praseodymium, cerium, and lanthanum, are recognized as strategic mineral resources both domestically and internationally [1]. These elements are widely used in neodymium–iron–boron (NdFeB) permanent magnets, cerium-based catalysts, and praseodymium–neodymium oxides (PrNdOx) and are globally recognized as strategic mineral resources [2]. However, as a non-renewable resource, the mining output of rare-earth elements can no longer meet the growing demand [3]. The stable supply of rare-earth elements in the future will mainly rely on their recovery from solid waste containing rare-earth elements.

At present, the recovery of rare-earth elements from solid waste mainly focuses on waste products such as used permanent magnets, wasted fluorescent materials, and abandoned polishing materials [4]. There is relatively little research on the rare-earth-containing waste slag generated during the preparation of rare-earth metals. Common preparation methods for rare-earth metals include molten-salt electrolysis, metallothermic reduction, vacuum distillation, etc., [5]. Among them, the molten-salt electrolysis method has the advantages of continuous production, low cost, high yield, and less waste slag, and is most widely used in actual production (accounting for more than 95%) [6]. According to statistics, in 2021, about 65,000 tons of rare earth were prepared by the molten-salt electrolysis method, and about 2500 tons of rare-earth molten-salt slag were generated. Rare-earth molten-salt slag contains a large amount of key and scarce rare-earth elements such as praseodymium (Pr), neodymium (Nd), and dysprosium (Dy) [7]. The efficient recovery and recycling of these rare-earth elements would not only diminish the necessity for extracting new rare-earth ores but also facilitate the establishment of a robust rare-earth recycling sector. Consequently, this would lead to a substantial enhancement in the overall efficiency and sustainability of rare-earth resource utilization.

Rare-earth molten-salt slag contains 10–80% of rare-earth oxide (REO), mainly existing in the forms of rare-earth fluorides and rare-earth oxyfluorides [8] and also contains a certain amount of lithium and fluorine elements. Currently, acid conversion, salt conversion, and alkali conversion methods are usually adopted to transform rare-earth fluorides into rare-earth compounds that are easily soluble in water or acid [9]. In the alkali conversion method, Qiu and others used the sodium hydroxide-roasting–acid-leaching process at 600 °C, which could make the leaching rate of rare-earth elements exceed 96% [10]. Lai and others carried out phase reconstruction by adding LiOH·H₂O for roasting at 600 °C and then vacuum distilling at 1100 °C, achieving a leaching rate of rare-earth elements of 99.27% [7]. In the salt conversion method, Tong adopted the co-roasting method of CaO and Al₂(SO₄)₃, which could convert rare-earth fluorides into soluble rare-earth sulfates at 900 °C [11]. Liang used the sodium silicate-roasting method at 850 °C to make the leaching rate of rare-earth elements reach 98.96% [5]. Wu adopted the sodium carbonate-roasting process at 700 °C to make the leaching rate of rare-earth elements exceed 98% [12]. In the acid conversion method, Tian used the sulfuric acid-leaching method to make the conversion rates of Nd, Pr, Dy, and Li reach above 95.00% at a temperature of 633 K [13]. Hu and Wang and others, respectively, used the nitric acid-roasting and sulfuric acid-roasting methods to convert rare-earth fluorides into soluble nitric acid/sulfuric acid rare earths at around 250 °C, with a conversion rate of rare-earth fluorides >95% [6,14]. Although the above processes can achieve a relatively high leaching rate of rare earths, they have high reaction temperatures, consume a large amount of energy, and are prone to causing secondary pollution. Against the backdrop of “carbon neutrality and carbon peaking”, it is urgent to develop more green and low-carbon treatment methods to achieve the efficient recovery of valuable elements in rare-earth molten-salt slag.

Compared with traditional alkali melting, roasting, and other processes, the sub-molten-salt medium with an alkali concentration greater than 50% has obvious thermodynamic advantages and superior kinetic performance [15]. It has a higher reaction efficiency in the extraction and refining of precious metals, light metals, and refractory metals and is particularly suitable for the recovery of valuable metals such as chromium, vanadium, aluminum, titanium, tantalum, and niobium [16]. In terms of vanadium extraction from vanadium slag, sub-molten-salt technology has achieved the synchronous and efficient extraction of vanadium/chromium resources in vanadium slag under atmospheric pressure and relatively low-temperature conditions. The recovery rates of vanadium and chromium reach more than 95% and 80%, respectively, while the vanadium extraction rate of the traditional roasting technology is only 80%, and chromium can hardly be recovered [17]. In the field of chromate preparation, with the sub-molten-salt-medium-enhanced dissolution technology under atmospheric pressure and at a temperature of 300 °C, the chromium recovery rate in chromite can reach 99%, which is more than 15% higher than that of the traditional roasting method [18]. Ma used the KOH sub-molten-salt medium to achieve the recovery of Al₂O₃ in low-grade diaspore bauxite, and the recovery rate reached 90.7% [19]. Zhou used the KOH sub-molten-salt method to extract niobium and tantalum from low-grade tantalum–niobium ores. Under optimal conditions, the leaching efficiency of niobium and tantalum were 98% and 96%, respectively [20]. Zhao used NaOH sub-molten-salt to leach silicon in laterite nickel ore under atmospheric pressure. Tong and others used the KOH sub-molten-salt method to prepare potassium titanate whiskers and titanium dioxide from high-titanium slag. Sun used the NaOH sub-molten-salt method to treat Al and Si in Bayer red mud, and all achieved good results [21].

To address the challenges of processing rare-earth molten-salt slag and promote green recycling, this study introduces an innovative atmospheric alkaline leaching methodology for resource recovery from molten-salt slag, rigorously optimizing process parameters to ensure robust results and conducting detailed kinetic analysis to deepen the understanding of the reaction mechanisms. The rigorously analyzed experimental data not only support the conclusions but also pave the way for future advancements in the field.

2. Materials and Methods

2.1. Raw Materials

The raw materials used in this experiment were obtained from a rare-earth metallurgical plant (Figure 1a). After being crushed and ground to 0.074 mm (Figure 1b), component and phase analyses were carried out. The main chemical components are shown in

Table 1. The chemical components of rare-earth molten-salt slag.

Component	Nd	Pr	F	Li	Ca	Al
Content (%)	43.58	12.39	33.23	3.47	0.91	0.21

. The rare-earth elements in the molten-salt slag are mainly Nd and Pr, with contents of 43.58% and 12.39%, respectively. The fluorine content is relatively high, reaching 33.23%, and the lithium content is 3.47%. Nd, Pr, F, and Li all have high recycling and utilization values.

The XRD analysis results of the sample are shown in Figure 2. It can be seen that the Nd, Pr, and Li in the slag mainly exist in the form of NdF₃, PrOF, and LiF.

2.2. Experimental Methods

First, weigh a certain amount of sodium hydroxide and place it in a three-necked flask. Then, add a proper quantity of water to prepare solutions for different concentrations and place a magnetic stirrer bar into the reactor, the experimental device was shown in Figure 3. Subsequently, heat the system to the set temperature. When the temperature stabilizes, add 10 g of molten-salt slag into the flask. After the reaction is completed, place the products in a 60 °C water bath and stir continuously for 15 min. Subsequently, filter the mixture to obtain the alkali-leaching solution. Dry the filter residue, then carry out the acid-leaching operation (3 mol/L HCl concentration, at 50 °C for 30 min, in a liquid-to-solid ratio of 10:1). After the acid-leaching reaction is finished, obtain the acid-leaching solution. Determine the contents of fluorine (F) and lithium (Li) in the alkali-leaching solution, as well as the content of Nd and Pr in the acid-leaching solution, respectively. The experimental flowchart is shown in Figure 4. Calculate the leaching efficiency according to the following formula:

$$\phi \left(\%\right)={\displaystyle \frac{V_1{C}_1+V_2{C}_2}{M_0{W}_0}}\times 100\%,$$

(1)

where φ (%) represents the leaching efficiency, %; V₁/V₂ is the volume of the alkali-leaching solution/acid-leaching solution, L; C₁/C₂ is the concentration of the component in the alkali-leaching solution/acid-leaching solution, g/L; M₀ is the mass of the reaction slag, g; W₀ is the initial mass fraction of each component in the reaction slag, wt.%.

2.3. Equipment and Reagents

The physical phases are analyzed using an X-ray diffractometer (XRD, Ultima IV, Rigaku, Tokyo, Japan). The contents of F and Li are determined using ion chromatography (IC; Thermo Scientific Integrion ICS-900, Waltham, MA, USA) and atomic absorption spectrometry (AAS; GGX-600, Haikuang, Taipei, China). The contents of Nd and Pr are determined using inductively coupled plasma emission spectroscopy (ICP; PerkinElmer 8300, Springfield, IL, USA). Sodium hydroxide (NaOH, analytical pure, XIHUA, Xihua, China) and hydrochloric acid (HCl, 36%, XIHUA, Xihua, China) are used as the main chemical reagents.

3. Results

3.1. Effect of Reaction Temperature on the Leaching Efficiency of Nd, Pr, F, and Li

Under the conditions of a sodium hydroxide mass concentration of 80%, a mass ratio of sodium hydroxide to molten-salt slag of 3:1, and a reaction duration of 120 min, the influence of different reaction temperatures (

Table 2. The experimental conditions and results on the effects of reaction temperature.

	Conditions				Leaching Efficiency (%)
	Alkali Concentration (%)	Alkali–Slag Mass Ratio	Time (min)	Temperature (°C)	Pr	Nd	F	Li
Ex1	80	3:1	120	120	42.16	44.11	32.55	28.22
Ex2	80	3:1	120	140	61.56	64.23	58.51	35.23
Ex3	80	3:1	120	160	89.78	91.23	90.715	72.72
Ex4	80	3:1	120	180	95.78	96.92	95.02	95.87
Ex5	80	3:1	120	200	96.12	96.98	97.58	95.88

) on the leaching efficiency of fluorine, lithium, and rare-earth elements was investigated. The experimental results are shown in Figure 5a and Table 2.

As can be seen from Figure 5a and Table 2, the leaching efficiency of Pr, Nd, F, and Li increases with the rise in the reaction temperature. When the reaction temperature increases from 100 °C to 180 °C, the leaching efficiency increases from 42.16%, 44.11%, 32.55%, and 28.22% to 95.78%, 96.92, 95.02%, and 95.87%, respectively. With the increase in the reaction temperature, the diffusion efficiency and chemical reaction rate increase accordingly, and the decomposition rate of the molten-salt slag is significantly improved. The decomposition efficiency of the molten-salt slag remains at a very low level before 140 °C, and lithium fluoride can hardly be converted into lithium hydroxide, remaining in the solid phase after washing with water. At 180 °C, the leaching efficiency of Pr, Nd, F, and Li is basically above 95%. When the temperature is further increased, the leaching efficiency basically tends to be stable, and there is no significant increase. Increasing the temperature has little promotion effect on the reaction. Therefore, 180 °C was selected as the experimental temperature for subsequent experimental exploration.

3.2. Effect of Reaction Time on the Leaching Efficiency of Nd, Pr, F, and Li

Under the conditions of a sodium hydroxide mass concentration of 80%, a mass ratio of sodium hydroxide to molten-salt slag of 3:1, and a reaction temperature of 180 °C, the influence of different reaction times (

Table 3. The experimental conditions and results on the effects of reaction time.

	Conditions				Leaching Efficiency (%)
	Alkali Concentration (%)	Alkali–Slag Mass Ratio	Time (min)	Temperature (°C)	Pr	Nd	F	Li
Ex6	80	3:1	10	180	35.23	36.78	32.22	27.45
Ex7	80	3:1	20	180	51.45	52.87	45.57	43.25
Ex8	80	3:1	30	180	60.05	61.28	58.22	52.77
Ex9	80	3:1	60	180	76.73	75.03	70.26	66.25
Ex10	80	3:1	90	180	87.12	88.35	83.45	75.55
Ex11	80	3:1	120	180	95.78	96.92	95.02	95.87
Ex12	80	3:1	150	180	97.55	97.72	95.22	96.25
Ex13	80	3:1	180	180	97.56	97.97	96.22	96.28

) on the leaching efficiency of fluorine, lithium, and rare-earth elements was investigated. The experimental results are shown in Figure 5b and Table 3.

As can be seen from Figure 5b and Table 3, the leaching efficiency of Pr, Nd, F, and Li in the slag increases rapidly at first and then gradually levels off as the reaction time prolongs. In the first 30 min of the reaction, the leaching efficiency increases significantly from 35.23%, 36.78%, 32.22%, and 27.45% to 60.05%, 61.28%, 58.22%, and 52.77%, respectively. However, when the reaction time increases from 30 min to 120 min, the increase in the leaching efficiency of Pr, Nd, F, and Li slows down, rising from 60.05%, 61.28%, 58.22%, and 52.77% to 95.78%, 96.92%, 95.02%, and 95.87%, respectively. Continuing to extend the reaction time results in little change in the leaching efficiency, with an increase of less than 1%. Further prolonging the reaction time has little significance for the decomposition of the molten-salt slag. Therefore, 120 min was chosen as the optimal reaction time.

3.3. Effect of the Alkali-to-Slag Ratio on the Leaching Efficiency of Nd, Pr, F, and Li

Under the conditions of a sodium hydroxide mass concentration of 80%, a reaction temperature of 160 °C, and a reaction time of 120 min, the influence of different alkali–slag ratios (

Table 4. The experimental conditions and results on the effects of alkali-to-slag ratio.

	Conditions				Leaching Efficiency (%)
	Alkali Concentration (%)	Alkali–Slag Mass Ratio	Time (min)	Temperature (°C)	Pr	Nd	F	Li
Ex14	80	1:1	120	180	49.78	51.23	48.68	15.45
Ex15	80	2:1	120	180	85.43	86.72	80.58	65.25
Ex16	80	3:1	120	180	95.78	96.92	95.02	95.87
Ex17	80	4:1	120	180	97.12	97.23	96.56	95.26
Ex18	80	5:1	120	180	96.12	96.38	94.35	93.25

) on the leaching efficiency of rare earths, fluorine, and lithium is shown in Figure 5c and Table 4.

It can be seen from the data in Figure 5c and Table 4 that as the alkali–slag ratio increases from 1:1 to 3:1, the leaching efficiency of Pr, Nd, F, and Li increase from 49.78%, 51.23%, 48.68%, and 15.45% to 95.78%, 96.92%, 95.02%, and 95.87%, respectively. In the case of a lower alkali–slag ratio, the decomposition efficiency of the molten-salt slag is low because, in this case, there is not enough sub-molten-salt medium to wrap the molten-salt slag. With the increase in the alkali–slag ratio, the leaching rate rises rapidly. A sufficient sub-molten-salt medium provides a sufficient liquid-phase environment, and a sufficient reaction medium can fully contact with mineral particles, so the reaction is accelerated. When the alkali–slag ratio exceeds 3:1, the leaching rate hardly changes much. Therefore, an alkali–slag ratio of 3:1 was selected for the subsequent experimental research.

3.4. Effect of NaOH Concentration on the Leaching Efficiency of Nd, Pr, F, and Li

Under a reaction temperature of 160 °C, a reaction time of 120 min, and an alkali–slag ratio of 3:1, the experimental results of the effect of different NaOH concentrations (

Table 5. The experimental conditions and results on the effects of NaOH concentration.

	Conditions				Leaching Efficiency (%)
	Alkali Concentration (%)	Alkali–Slag Mass Ratio	Time (min)	Temperature (°C)	Pr	Nd	F	Li
Ex19	50	3:1	120	180	74.78	75.97	72.56	50.25
Ex20	60	3:1	120	180	84.33	86.55	83.25	75.89
Ex21	70	3:1	120	180	89.18	91.04	91.45	85.25
Ex22	80	3:1	120	180	95.78	96.92	95.02	95.87
Ex23	90	3:1	120	180	89.34	90.54	89.45	75.65

) on the rare-earth leaching rate are shown in Figure 5d.

According to the data in Figure 5d and Table 5, when the NaOH concentration is increased from 50% to 80%, the leaching efficiency of rare earths, fluorine, and lithium increase from 74.78%, 75.97%, 72.56%, and 50.25% to 95.78%, 96.92%, 95.02%, and 95.87%, respectively. When the NaOH concentration is further increased to 90%, the leaching rate will decrease. According to the reaction phenomenon, the decomposition of the molten-salt slag mainly occurs in the early stage. At this time, the slurry has good fluidity, the molten-salt slag can quickly make contact with the medium during stirring, and the decomposition rate of the molten-salt slag increases rapidly. In the case of a low concentration, the ionic activity of the system is not high. Although the decomposition effect of the molten-salt slag is good in the early stage of the reaction, the overall decomposition efficiency is lower than that in the high-concentration case. When the sub-molten-salt medium concentration is too high, the fluidity of the slurry is reduced, which affects the contact between the molten-salt and the reaction medium. Therefore, a NaOH concentration of 80% was selected for subsequent experimental research.

3.5. Comparison of Energy Consumption and Environmental Benefits Between This Experiment and Current Common Methods

Compared to the alkali-fusion method, the salt-roasting method, and the acid-leaching method, the approach employed in this experiment demonstrates significant advantages in terms of energy consumption, economic feasibility, and environmental protection (

Table 6. Comparison between this experiment and current common methods [9].

Method	Temperature	Reagent	Energy Consumption	Recovery Rate (REEs)	Product Purity	Solid/Gas Water Waste
Alkali-fusion method	950 °C	Ca(OH)2	High	97%	Low	Yes
	600 °C	NaOH	High	99%	High	Yes
	600 °C	LiOH·H2O	High	99.27	High	No
	600 °C	NaOH	High	96%	High	Yes
	300–500 °C	NaOH	High	97%	Low	Yes
Salt-roasting method	700 °C	Na2B4O7·10H2O	High	97%	Low	Yes
	630 °C	CaO, Al2(SO4)3	High	90%	High	Yes
	850 °C	Na2SiO3	High	98.96%	High	No
	700 °C	Na2CO3	High	99.13%	Low	Yes
	750 °C	(NH4)2SO4	High	95%	High	Yes
	700 °C	Na2CO3	High	98%	High	No
Acid-leaching method	333 °C	H2SO4	Low	95%	High	Yes
	250 °C	HNO3	Low	98%	High	Yes
This experiment	180 °C	NaOH	Low	96.35%	High	No

). The alkali-fusion and salt-roasting methods necessitate high temperatures (ranging from 300 °C to 950 °C), leading to substantial energy consumption, increased operational costs, and a larger carbon footprint. These methods also generate gaseous emissions, posing environmental challenges. In contrast, the acid-leaching method and the method utilized in this experiment operate at lower temperatures (333 °C and 180 °C, respectively), thereby reducing energy consumption and operational costs. However, the acid-leaching method produces significant acidic emissions, contributing to environmental pollution. The method employed in this experiment not only operates at a lower reaction temperature and does not generate gaseous emissions, thereby minimizing the environmental impact, reducing harmful emissions, and providing a more sustainable and economically viable solution for the recovery of rare-earth elements. Consequently, the method adopted in this experiment represents a superior approach in terms of energy efficiency, economic feasibility, and environmental sustainability.

3.6. Mineral Phase Transformation and Morphological Analysis

Under the optimal experimental conditions, the product obtained after reaction and subsequent water washing was subjected to laser particle size analysis. The analysis results, as depicted in Figure 6, reveal that the particle size distribution of the post-reaction product exhibits an unimodal pattern and approximates a normal distribution. This indicates that the particle size of the product is relatively homogeneous. Notably, in comparison with the original ore, the average particle size of the product was substantially reduced, dropping from 48.92 μm to 1.1623 μm.

By conducting a comparative phase analysis between the original ore (shown in Figure 1) and the reactant under the optimized conditions (Figure 7b), it was ascertained that the phases of rare-earth fluorides, rare-earth oxyfluorides, and lithium fluoride within the rare-earth molten-salt electrolysis slag remained undetected. Instead, the predominant constituent in the reaction product was rare-earth hydroxide. From the morphological analysis presented in Figure 7a, it can be discerned that the phase reconstruction process of rare-earth fluoride initiates from the surface and culminates in the decomposition into small crystalline particles. Through energy-dispersive spectroscopy (EDS) analysis (Figure 7A,B), the elements identified at the selected point A, namely Nd, Pr, O, and F, correspond to the unreacted raw materials (molten-salt electrolysis slag) in the inner layer. At the selected point B, the detected elements are Nd, Pr, and O. Although the H element cannot be directly detected by EDS, in conjunction with XRD analysis, it can be conclusively determined that the substance is RE(OH)₃. In the figure, certain reaction products display clear outlines (in a strip-like form) and adhere to the surface of the raw material, while others possess indistinct outlines, suggesting that the raw material has undergone corrosion, thereby leading to its gradual diminution. In light of the above comprehensive analysis, the absence of lithium and fluorine in the outer-layer particles implies that, via the water-washing process, sodium fluoride and lithium hydroxide, which possess relatively high solubilities, have migrated into the liquid phase. The reacted products manifest on the surface of the molten-salt slag, and the unreacted rare-earth molten-salt slag particles progressively shrink, conforming to the shrinking core model.

4. The Kinetic Model of Leaching Reaction

4.1. The Controlling Step and Apparent Activation Energy of Fluorine Element Conversion

The relationship between the leaching rate of element F and reaction time under diverse temperature conditions is illustrated in Figure 8a. The leaching rate data at 180 °C were, respectively, fitted with three simplified kinetic equations, and the outcomes are presented in Figure 8b–d. As can be discerned from the fitting data, among the fittings of the three kinetic models, the model 1 − (1 − x)^(1/3) exhibits the highest fitting degree, with R² attaining 0.9969. In contrast, the R² values of the other two kinetic models, 1 − (2/3)x − (1 − x)^(2/3) and 1/3ln(1 − x) + (1 − x)^−(1/3), are 0.9868 and 0.8370, respectively. The fitting degree of the internal diffusion model is somewhat lower than that of the interfacial chemical model, and the fitting degree of the mixed reaction process model integrating internal diffusion and chemical reaction is rather poor. Consequently, the controlling step in the conversion process of fluorine element is presumed to follow the reaction control model.

The variation pattern of the leaching rate of fluorine element with time under different temperatures was fitted using the model 1 − (1 − x)^(1/3), yielding Figure 8e. Subsequently, the slopes of the curves obtained at various temperatures in Figure 8e were fitted against 1000/T, generating Figure 8f. Based on the slope, the apparent activation energy of the leaching process of element F was computed to be 59.06 kJ/mol, which exceeds 40 kJ/mol. Both the fitting degree and the apparent activation energy data corroborate that this leaching process adheres to the chemical reaction control model.

4.2. The Controlling Step and Apparent Activation Energy of Li Conversion

The relationship between the leaching rate of the Li element and the reaction time under different temperature conditions is shown in Figure 9a. The leaching rate data at 180 °C were fitted with three simplified kinetic equations, respectively, and the results are shown in Figure 9b–d. It can be seen from the fitting data that among the fittings of the three kinetic models, the model 1 − (1 − x)^(1/3) has the highest fitting degree, with R² = 0.9962, while the R² values of the other two kinetic models, 1 − (2/3)x − (1 − x)^(2/3), 1/3ln(1 − x) + (1 − x)^−(1/3), are 0.9943 and 0.8466, respectively. The fitting degree of the internal diffusion model is slightly lower than that of the interfacial chemical model, and the fitting degree of the mixed reaction process model of the internal diffusion–chemical reaction is relatively poor. Therefore, the controlling step of the Li element conversion process is speculated to be the reaction control model.

The variation law of the leaching rate of fluorine element with time at different temperatures was fitted with the model 1 − (1 − x)^(1/3), and Figure 9e was obtained. The slopes of the curves obtained at different temperatures in Figure 9e were fitted with 1000/T, and Figure 9f was obtained. According to the slope, the apparent activation energy of the F element leaching process was calculated to be 57.33 kJ/mol, which is greater than 40 kJ/mol. Both the fitting degree and the apparent activation energy data indicate that this leaching process conforms to the chemical reaction control model.

5. Conclusions

The maximum leaching efficiency of Nd, Pr, F, and Li can be reached at 96.92%, 95.78%, 95.02%, and 95.87% under the optimized conditions of an NaOH concentration of 80%, a reaction temperature of 180 °C, a reaction time of 2 h, and an alkali–slag mass ratio of 3:1. The mineral phase analysis indicates that the rare-earth fluorides/oxyfluorides transformed into rare-earth hydroxide, while F and Li were converted into soluble NaF and LiOH.

The morphology analysis indicates that the reaction initiated from the surface of the raw material, forming small crystalline particles, accompanied by corrosion of the inner layer, leading to gradual shrinkage of the unreacted core, consistent with the shrinking core model. The leaching process of F and Li are controlled by interfacial chemical reactions, and their apparent activation energies are 57.33 kJ/mol and 40 kJ/mol, respectively.