Hydrometallurgical Process to Extract Niobium from Tin Slag Through Alkaline Treatment and Sulfuric Acid Leaching

Abstract

Niobium and tantalum are critical metals that are important for technological development. Their main applications are in the production of alloys for the civil construction, electronics, nuclear, and aerospace industries, and in catalysis. Tin reduction slag is a possible secondary source of niobium and tantalum, containing 3.7% and 0.5% of Nb and Ta, respectively. The slag matrix is mainly composed of calcium silicate, a low-reactivity material that prevents contact between the leaching solution and the metals to be extracted; therefore, it is necessary to previously react the material with molten NaOH. This reaction converts calcium silicates into sodium silicates, which are more reactive and water-soluble, and converts the metals into oxyanions, niobates, and tantalates, which are more reactive species. After treatment with molten hydroxide, the material is then solubilized in water; this reaction removes part of the soluble materials and also fragments the silicate matrix. Nb and Ta remain in the solid phase during the water washing step and then undergo acid leaching, where, after the parameters are evaluated, Nb extraction of 96% and Ta leaching of less than 3% are achieved, using a concentration of 10 mol/L H₂SO₄, a time of 2 h, a temperature of 90 °C, and a liquid–solid ratio of 50.

1. Introduction

Niobium and tantalum are two transition elements belonging to group 5 of the periodic table, and have much in common. Due to their similarity in electronic distribution, these elements exhibit comparable physical and chemical properties, which favor their joint association in nature and their application in specific technologies that are essential for technological development.

Among the main applications of niobium is the production of high-strength low-alloy (HSLA) steels, which are widely used in construction, oil and gas pipelines, modern automobiles, and stainless steels [1]. Other emerging applications include its use in the production of superconducting magnets for MRI (magnetic resonance imaging) machines [1]. In the materials industry, pure niobium oxide is used in optical applications [2]. Finally, the metal has been studied as a component of the anode in lithium-ion batteries to improve their performance [3].

Similarly to niobium, tantalum is also applied in the production of alloys, with particular emphasis on corrosion-resistant alloys [4]. However, its main application is the manufacture of capacitors in the electronics industry. A mixture of metallic tantalum powders and tantalum oxide is used to form the cores of miniaturized capacitors, which are widely applied in electronic devices such as computers and mobile phones [1].

According to the annual survey conducted by the United States Geological Survey (USGS) through the National Minerals Information Center, the main declared global reserves of niobium are located in Brazil (94%) and Canada (~6%), with more than 90% of global production concentrated in Brazil. The report further states that most of the economically exploitable material occurs in the form of pyrochlore in carbonatite (igneous rocks that contain more than 50 vol.% carbonate minerals) [5].

In contrast, tantalum reserves are more concentrated in the Democratic Republic of the Congo (61%) and Brazil (28%), with the highest production occurring in the Congo itself and other sub-Saharan African countries, which together account for a total of 77% of the declared global production [6].

Due to their economic importance, the difficulty in substituting them in certain applications, and their geographic concentration in a few regions of the globe, both metals are considered critical by the main economic blocs worldwide. The European Union has classified both metals as critical since 2017, highlighting precisely their poor global distribution and the importance of these metals in technological applications [7]. The National Research Council of the United States, in addition to considering both metals to be critical, emphasizes that tantalum is a conflict metal, as most of its production occurs in regions of the African continent where conflicts exist due to mining activities [8].

With most reserves concentrated in South America and Africa, these metals are considered extremely critical for Asian countries such as China and Japan. As a result, strong trade relations and investments in metal-producing countries are common strategies adopted by these economic powers. In research conducted in 2023 in Japan, niobium was identified as the element with the highest criticality in terms of supply risk, while tantalum ranked fourth [9]. For China, the Ministry of Industry and Information Technology considers niobium one of the most critical elements due to its importance in the construction industry and technological applications [10]. The country’s main response has been investment in producing countries such as Brazil. As an example, the CMOC (China Molybdenum Corporation), the second-largest niobium producer worldwide, is a Chinese company that extracts and refines this metal in Brazil [3].

For metal-producing countries such as Brazil, these metals are considered strategic, and there are incentives for research not only aimed at improving metal extraction and production processes, but also at developing new applications for them [11].

The traditional processes for the extraction and production of Nb and Ta products can be divided into two groups: those aimed at producing metals for alloy manufacturing, and routes focused on the recovery of high-purity materials for applications in the electronics and optical industries. In alloy production, the process generally consists of ore extraction and concentration, followed by carbothermic reduction processes for impurity removal and aluminothermic reduction for metal production. Although this route is efficient and widely employed, it does not effectively separate Nb and Ta, and it also presents difficulties in separating elements such as Th and U, which may end up in the final products, resulting in metallic alloys composed of Fe, Nb, Ta, and associated contaminants [12]. Alternatively, hydrometallurgical processes can be used, in which the metals are solubilized in aqueous solutions, generally using potassium hydroxide or acidic mixtures containing HF. The advantage of this approach is the possibility of separating the metals through techniques such as solvent extraction once they are in solution. The disadvantages include the generation of large volumes of effluents concentrated in environmentally hazardous ions, such as hydroxides and fluorides [12].

As an alternative to this type of process, several authors have sought alternatives, such as the use of fluoride salts that generate HF in situ, thereby reducing the hazards associated with handling this reagent [13], as well as studies investigating different mixtures of alkaline compounds, such as sodium carbonates and hydroxides, to promote the extraction of these metals into solution [14].

As important as studying and improving the metal extraction process itself is the investigation of possible secondary sources for these metals. As discussed, the primary production of these metals is concentrated in a few, politically unstable countries, which contributes to their criticality. Therefore, the study of secondary sources, especially mining residues and discarded capacitors, has been increasingly explored as a potential source for the recovery of these metals [11].

The main secondary source studied for Nb recovery is tin slag, as both metals commonly occur together in nature, and during the tin refining process, niobium remains in an oxidized form in the slag [15]. Flotation residues from the same industry also contain significant amounts of these metals and can therefore be investigated as secondary sources [16]. Associations of Nb and Ta with titanium have also been reported, and in mining residues of this metal, it is also possible to find significant quantities of Nb and Ta, making them promising secondary sources for the recovery of these metals [17]. The recovery of Nb and Ta from tin slag has been studied using different approaches (

Table 1. Strategies for Nb and Ta recovery or concentration in tin slag.

Strategy	Procedure	Results	Reference
Chlorination and Carbochlorination	Using gas mixtures (Cl2 + N2 and Cl2 + CO + N2) at high temperatures (~1000 °C) to form soluble chlorides and carbonates	Leaching: 84% Nb 65% Ta	[18]
Direct leaching with fluoride salt (NH4F-HCl)	Direct leaching of the ground slag using a mixture of HCl and NH4F to generate HF in situ for the complexation of Nb and Ta	Leaching: 100% Nb 5% Ta	[13]
A-B-A (acid–base–acid) or B-A-B (base–acid–base) A: HCl and HF B: NaOH	Sequential acid–base–acid leaching selectively dissolves matrix, concentrating tantalum-niobium oxides.	Enrichment of the slag from 15 wt.% (Nb + Ta oxides) to 63 wt.%, with a loss of 14% of the original Nb and Ta mass.	[17]
NaOH treatment and H3PO4 leaching	Reaction with NaOH (8 M) and H3PO4 (0.5–1.5 M) for Nb–Ta enrichment	Tantalum enrichment from 0.23% to 0.85% and niobium enrichment from 0.47% to 1.45%	[19]
NaOH treatment and HCl Leaching	Alkali roasting with NaOH removes impurities, followed by HCl acid leaching	Residue enriched threefold, reaching ~10% Nb2O5 and Ta2O5	[20]

), with some studies predicting the concentration of the metals in the solid residue due to the difficulty in leaching these refractory metals.

When comparing the studies, it is understood that, for leaching without the use of complexing agents such as HF, acidic or alkaline thermal treatments are required to decompose the slag matrix and form soluble Nb and Ta phases; even after these processes, leaching of the refractory metals remains difficult, requiring high acid concentrations.

Within this context, this work aims to study the use of alkaline treatment with sodium hydroxide, followed by water and sulfuric acid leaching, as an alternative route for the solubilization of Nb and Ta from a tin slag rich in calcium silicate.

2. Materials and Methods

The tin slag was received from a tin company in Brazil, representing the final residue produced in the cassiterite beneficiation. The slag was dried, homogenized and quartered. The sample was milled until <75 µm to achieve an increase in the superficial area and matrix liberation [21], and a representative part was characterized.

2.1. Characterization

The tin slag was characterized through different techniques. The mineralogical composition was determined by X-ray diffraction (XRD, Miniflex 300, Rigaku, The Woodlands, TX, USA), the scanning was performed over a 2θ range of 20–80°, with a speed of 1.5°/min and a step size of 0.02°. The morphological analysis was carried out in a scanning electron microscope with energy-dispersive spectroscopy (SEM-EDS, JCM-7000, Jeol, Tokyo, Japan) with a voltage of 15 kV and backscattering electrons. The sample was also kept in a vacuum chamber for 30 min to eliminate any residual moisture and volatiles that could damage the equipment.

A granulometric analysis was also carried out before milling to evaluate if Nb and Ta were concentrated in a certain range of particle sizes, predicting the physical concentration. Sieves with mesh sizes of 10 to 400 were used, and the solids were digested for Nb and Ta determination.

The chemical analysis was performed via energy-dispersive X-ray fluorescence (EDXRF, Epsilon 3, Malvern Panalytical, Worcestershire, UK) to obtain a qualitative result, and via atomic absorption spectroscopy (AAS, AA-7000, Shimadzu, Kyoto, Japan) and inductively coupled plasma optical emission spectrometry (ICP-OES, 710 series, Agilent, Santa Clara, CA, USA) to obtain a quantitative result. The digestion of the slag was performed by fusion with lithium metaborate in a proportion of 1:6 (kg slag/kg LiBO₂) in a furnace at 1000 °C for 30 min, followed by acid digestion in HCl 10%.

2.2. Experimental Procedure

The methodology for Nb and Ta extraction from tin slag was divided into 3 parts: alkaline treatment, water leaching, and acid leaching. The purpose of the alkaline treatment was to solubilize part of the Si present in the slag by converting the CaSiO₃ species into soluble species like Na₄SiO₄ (Equation (1)) [22].

4CaSiO₃ + 12NaOH = 4Ca(OH)₂ + 3Na₄SiO₄ + 2H₂O

(1)

The CaSiO₃ structure confined the Nb and Ta species, interfering with the extraction rate. After the conversion to Na₄SiO₄, this species was then removed in the water leaching phase, making the species available. The Nb and Ta species also reacted with the NaOH (Equations (2) and (3)), producing more leachable reactive species [23].

3Nb₂O₅·nH₂O + xNaOH = Na_xH_8−xNb₆O₁₉·mH₂O + (3n − m + x − 4)H₂O

(2)

3Ta₂O₅·nH₂O + xNaOH = Na_xH_8−xTa₆O₁₉·mH₂O + (3n − m + x − 4)H₂O

(3)

The alkaline treatment was performed by adding the milled sample and the NaOH to a crucible, which was then placed in the furnace at 700 °C for 3 h. Different ratios between slag and NaOH were studied. The conditions for the alkaline treatment are displayed in

Table 2. Alkaline treatment conditions.

Parameters	Conditions
Time (h)	3
Temperature (°C)	700
Ratio (g NaOH/g slag)	0.5:1, 1:1, 2:1

.

After cooling, the solid material was disaggregated, and the mixture was leached with ultrapure water in a solid–liquid ratio of 1:50 (g/mL), at 60 °C, for 30 min. The solid was then filtered and dried at 100 °C for 24 h. The treated sample was then leached with H₂SO₄.

The acid leaching experiments were carried out in a reactor under magnetic stirring (IKA, Staufen im Breisgau, Germany) with a condenser system to avoid acid losses. All the experiments were carried out at 90 °C, and the acid concentration, the time, and the ratio between sample and acid were studied, with the conditions as displayed in

Table 3. Leaching experiment conditions.

Parameters	Conditions
Temperature (°C)	90
Acid concentration (mol/L)	3, 6, 7, 10, 18
Time (h)	2, 4, 6, 18
Ratio (g sample/mL acid solution)	1:10, 1:20, 1:30, 1:50

.

The first condition evaluated was the acid concentration, with the parameters set to 18 h and 1:50 (g sample/mL acid solution). Then, the time was evaluated with the best condition for concentration and the same ratio between sample and acid (1:50). Finally, the ratio was evaluated as the best condition for concentration and time.

After the determined time, the leaching solution was filtered and analyzed by ICP-OES to determine the Nb and Ta extraction rate. The solution obtained with the higher extraction was also analyzed to determine the chemical composition, predicting future purification processes.

3. Results and Discussion

3.1. Characterization

The XRD analysis of the tin slag identified three crystalline phases: ZrO₂ (00-086-1451), SiO₂ (00-082-1563), and CaOSiO₂ (00-076-0186) (Figure 1). These phases are expected, given the silicon present in the soil in which cassiterite is found, the calcium utilized in the slagging process, and the zirconium also present in Amazonian soils [3,13,14].

The granulometric analysis demonstrated that the sample was concentrated in the 120–35 mesh window, which corresponds with 0.125–0.5 mm (

Table 4. Granulometric analysis with Nb and Ta contents.

Mesh	Sample Fraction (%)	Nb Content (%)	Ta Content (%)
10	0.6	3.9	0.6
18	1.2	3.6	0.6
35	20.2	3.6	0.6
60	29.3	3.6	0.6
120	23.6	3.5	0.6
200	10.8	3.6	0.6
400	9.7	3.6	0.6
<400	4.6	3.9	0.6

). However, the concentration of Nb and Ta was not observed at any size, with the Nb content being around 3.6% and the Ta content 0.6% at all sizes evaluated. Therefore, a possible physical concentration before the extraction process was discarded.

The SEM–EDS analysis indicated that the slag is mainly composed of two types of particles, with one of them being unreacted silica (Figure 2). This can be attributed to an excess of silica in the slagging process or to contamination occurring after the smelting process. Such contamination is common, since this material is comminuted and deposited in open stockpiles after being removed from the furnace.

The second type of particle corresponds to the slag itself, formed of two main components: a matrix composed of calcium silicate, indicated by the blue color in Figure 3, and, embedded within this matrix, zirconium oxide particles and dendrites containing niobium, iron, manganese, and thorium, indicated by the red color in Figure 3.

The EDXRF qualitative analysis indicated the presence of the following metals in the tin slag: Al, Ca, Fe, K, Mg, Mn, Na, Nb, Si, Sn, Ta, Th, Ti, U, Y, Zn, and Zr. Therefore, these elements were quantified by AAS (Si) and ICP-OES (other metals). The complete chemical analysis (

Table 5. Chemical analysis of the tin slag.

Al	Ca	Fe	K	Mg	Mn	Na	Nb	Si
2.6%	9.0%	3.5%	1.2%	1.8%	0.4%	0.2%	3.7%	19.9%
Sn	Ta	Th	Ti	U	Y	Zn	Zr	O (balance)
1.9%	0.5%	1.7%	1.1%	0.4%	0.2%	0.4%	11.9%	39.6%

) indicated that the elements with the highest concentrations presented in the slag are Ca (9.0%), Si (19.9%), and Zr (11.9%). The elements of interest in this study, Nb and Ta, were quantified at 4.6% and 0.5%, respectively. Other impurities that could interfere in the leaching process were also quantified, such as Al (2.6%) and Fe (3.5%).

These findings can be compared with those obtained by Garjulli et al. [15], Anes et al. [13] and Machaca et al. [3]. They also corroborate the XRD analysis, which demonstrated phases containing Zr, Si, and Ca.

3.2. Nb and Ta Extraction

The extraction of Nb and Ta reached a maximum of 48% and 41%, respectively, in the alkaline treatment study (Figure 4), when using a ratio of 1 kg NaOH/kg slag. At a solid–solid ratio of 0.5 kg NaOH/kg of slag, the extractions were 32% for Nb and 7% for Ta, and at the ratio of 2 kg NaOH/kg of slag, the extractions reached 23% for Nb and 12% for Ta. The extraction was quantified after the water and acid leaching, for which the conditions were: 30 min, 60 °C and 1:50 (g/mL) for the water leaching, and H₂SO₄ 3 mol/L, 18 h, 90 °C and 1:50 (g/mL) for the acid leaching.

The reduction in the extraction of Nb and Ta with the increase in the solid–solid ratio can be explained by the silica polymerization, which made it difficult to filter the leaching solution. The solution was imprisoned in the gel formed, therefore interfering with the extraction quantification. The polymerization of silica occurs with soluble Si in acid media, with the Si species connecting with each other, entrapping the solution inside [24].

With the rising of the NaOH/slag ratio, more Si was available in solution to react with the acid, due to the hydroxide reaction. In acidic pH (<7), Si is found in the form of silicic acid (Si(OH)₄), which reacts with the same species (Figure 5), resulting in polysilicic acid. The polysilicic acid species react with each other, forming a gel. This gel traps the solution, interfering with the filtration process and therefore misleading the extraction quantification [24,25].

The XRD analysis of the slag during alkaline treatment demonstrates the formation of a Na₂Si₂O₅ phase, which is soluble in water (Equation (4)) [26]. When comparing the diffractograms of the slag before (Figure 6A) and after the alkaline treatment (Figure 6B), it can be observed that the SiO₂ phase vanishes, and the CaSiO₃ phase has a decreased intensity due to the conversion to the Na₂Si₂O₅ phase.

Na₂Si₂O₅ + H₂O = 4Na⁺ + HSiO₄³⁻ + OH⁻

(4)

An XRD analysis after the water leaching (Figure 6C) was also performed to demonstrate the disappearance of the soluble Na₂Si₂O₅ phase. The alkaline treatment and the water leaching combined removed around 30% of the Si present in the slag. There was also 2% solubilization of Nb, and solubilization of Ta was not observed.

When studying the acid concentration effect on Nb and Ta extraction, an increase in Nb extraction was observed from 3 mol/L (48% extraction) to 18 mol/L (93%) (Figure 7). Given that Nb extraction was the same in 10 mol/L and 18 mol/L, 10 mol/L was the concentration utilized in the following studies. The increase in Nb extraction can be explained by the increase in reactant available.

However, Ta extraction decreased with the increase in concentration from 3 mol/L (41%) to 18 mol/L (1%). At higher acid concentrations, Ta can react and form a passivated layer of insoluble Ta₂O₅, which prevents leaching.

In the study by Akli et al. [19], the sodium tantalate formed during alkaline treatment was converted into tantalum pentoxide upon contact with phosphoric acid (Equation (5)). We believe that a similar phenomenon may have occurred with sulfuric acid in the present work, with increasing acid concentration favoring this process. Since tantalum pentoxide is insoluble in acidic media, its formation can be considered a passivation process, thereby reducing metal leaching.

3NaTaO₃ + H₃PO₄ = 1.5Ta₂O₅ + Na₃PO₄ + 1.5H₂O

(5)

Regarding reaction time, periods from 2 to 18 h were evaluated (Figure 8); for both metals, no variation in extraction was observed. Nb remained constant above 95%, while Ta leaching was negligible, lower than 3% in all experiments. This indicates that the process limitations are not kinetic but instead depend on other factors.

With the reduction in time from 18 to 2 h, the liquid-to-solid ratio, that is, the proportion between the leaching agent used and the treated slag, was then evaluated. Although ratios of up to 100 are reported in the literature [11], that is, 1 g of solid material to 100 mL of leaching agent, an attempt was made to investigate whether it would be possible to reduce this value, since the solid–liquid ratio is a critical parameter for hydrometallurgical systems [27].

Although the results obtained with a liquid-to-solid ratio of 1:30 were similar to those obtained at 1:50 (Figure 9), lower ratios show significant decreases in Nb extraction, reaching values 30 to 50 percentage points lower. This is attributed to the low solubility of the possible products formed in the leaching reactions. It is also possible to observe that, regardless of the amount of leaching agent or reaction time, tantalum leaching is always practically null, reinforcing the inertness of this metal and its compounds under the studied conditions.

Among the studied parameters, the best process conditions were a sulfuric acid concentration of 10 mol/L, a liquid-to-solid ratio of 1:50, a temperature of 90 °C, and a reaction time of 2 h, achieving 96% Nb leaching and less than 3% Ta leaching. The leaching solution concentrations for future purification studies are also displayed in

Table 6. Chemical analysis of the final leaching solution (mg/L). Alkaline treatment conditions: 3 h, 700 °C and 1:1 (g/g). Water leaching conditions: 30 min, 60 °C and 1:50 (g/mL). Acid leaching conditions: H₂SO₄ 3 mol/L, 2 h, 90 °C and 1:30 (g/mL).

Al	Fe	K	Mg	Mn	Nb	Ta
713	1121.1	75.8	807	115.5	1284.6	2.6
Th	Ti	U	Y	Zn	Zr
608.4	373.5	162	22.9	141.1	3095.9

.

Comparing this work with the studies by Akli et al. [19] and Bandopadhyay et al. [20], which also employed similar strategies (alkaline thermal treatment with NaOH followed by acid leaching), the results of the present study are extremely positive. Unlike previous works, which mainly focused on concentrating Nb and Ta in the solid residue due to the difficulty of leaching these metals, this study demonstrates that Nb can be effectively leached into solution. This process enables subsequent purification steps, such as solvent extraction and ion exchange. Moreover, the absence of fluoride species makes the process environmentally favorable and operationally safer.

4. Conclusions

This work concludes, as expected, that a possible route for Nb extraction from tin slags is sulfuric leaching carried out after reaction with molten sodium hydroxide and water leaching. In this process, the hydroxide reaction is fundamental because it converts the calcium silicate matrix into sodium silicate, which is more soluble, and converts Nb and Ta into more reactive species. The ideal ratio between slag and hydroxide evaluated in this work was 1:1, at 700 °C, for 3 h. After this process, the material is then leached in hot water (60 °C) for 30 min, at a liquid–solid ratio of 1:50 (g/mL), so that the silicate is solubilized and the niobates are released to react with the acid. Finally, the acidic step (sulfuric acid concentration of 10 mol/L, liquid-to-solid ratio of 1:50 (g/mL), temperature of 90 °C, and reaction time of 2 h) solubilized 96% of the Nb in the material.