Antimony Recovery from Industrial Residues—Emphasis on Leaching: A Review

Abstract

Antimony (Sb) is a metalloid widely used in different areas—from the cutting-edge renewable energy technologies to “classical” lead acid batteries. Its availability in primary sources is limited, and these sources are geographically unevenly distributed worldwide. Antimony use will increase in the future. That is why Sb is included in the critical raw material lists of the European Union and the USA. In order to mitigate the future Sb shortage, Sb recovery from industrial residues is worth considering. This paper presents the availability of Sb in nonferrous metals extraction waste and the applicability of the hydrometallurgical route for Sb recovery from such sources. Leaching is emphasized. The use of acidic and alkaline leaching methods, their recent modifications, and the effect of different process parameters (reagents’ type, solid-to-liquid ratio, temperature, and the addition of oxidizing reagents) are highlighted. The use of new leaching systems, such as deep eutectic solvents and non-aqueous solutions, is presented. Initial attempts to apply bioleaching are described. Finally, some proposals for future investigations are given.

1. Introduction

Antimony (Sb) is a silvery-white metalloid. Sb abundance in the Earth’s crust is accessed to be 0.2 mg/kg. It is a chalcophile element that is found in the sulfur (S)-containing minerals (sulfides and sulfosalts) of lead (Pb), copper (Cu), and silver (Ag). Stibnite (Sb₂S₃) is the predominant Sb mineral, which is found in nature among the over 100 existing Sb minerals. Stibnite is the mineral that is most important from the industrial point of view. Other S-containing Sb-bearing minerals of industrial interest are jamesonite (Pb₄FeSb₆S₁₄), tetrahedrite (Cu₁₂Sb₄S₁₃), zinkenite (Pb₉Sb₂₂S₄₂), boulangerite (Pb₅Sb₄S₁₁), bindheimite (Pb₂Sb₂O₆(O,OH)), berthierite (FeSb₂S₄), livingstonite (HgSb₄S₇), etc. Valentinite (rhombohedral Sb₂O₃), senarmontite (cubic Sb₂O₃), cervantite (Sb₂O₄ = Sb₂O₃·Sb₂O₅), stibiconite (H₂Sb₂O₅ = Sb₂O₄·H₂O), and kermesite (Sb₂S₂O) can be found among oxide ores containing Sb and of industrial interest [1,2]. The processing of Sb-containing Pb and gold (Au) ores is another source of primary Sb [3].

Antimony trioxide (Sb₂O₃) is the most important compound of Sb that is produced [1,4]. It is widely used as a flame retardant in plastics, rubbers, and coatings—in electronics, textiles, paper, etc.—due to its ability to decrease the amount of needed halogenated flame retardants. Antimony and its compounds are applied as a catalyst and stabilizer in the chemical industry and as an additive in ceramic and glass products. For many years one of the main applications of Sb is to produce Pb-Sb alloys that due to their high mechanical strength are used in automotive lead–acid batteries and in construction [2,5,6,7].

As an example, in the European Union (EU), Sb is used in the following sectors: flame retardants—43%, lead–acid batteries—32%, Pb alloys—14%, plastics (catalysts and stabilizers)—6%, and glass and ceramics—5% [8]. The major areas of Sb use in the United States are metal products, including antimonial lead and ammunition—43%, flame retardants—35%, and nonmetal products, including ceramics, glass, and rubber products—22% [9].

Antimony is used in some fast developing and emerging technologies, such as (i) for the development and application of diodes and IR detectors, which are useful in the environment and for working place monitoring, medical imaging, night vision, and military applications [10,11,12], and (ii) for the development of different non-volatile memory types—used in computer science and industrial automation [2].

Antimony and its compound are gathering large attention as materials suitable for application in renewable energy sources, such as solar cells, including thin-film and third-generation solar cells [13,14,15,16]. Antimony-based anodes have been proposed for lithium-ion batteries due to their good theoretical capacity, high electronic conductivities, high volumetric capacity, and high reversible lithium storage capacity [17,18]. Due to its high theoretical capacity (946 mA h g⁻¹) and better ability to intercalate sodium ions, compared to different types of graphite, Sb₂S₃ is one of the most studied and proposed materials for anodes in sodium-ion batteries [19,20].

From its primary resources Sb is extracted by the pyrometallurgical and hydrometallurgical routes. The share of pyrometallurgy is still very high (90–95%) [2,21]. Stibnite concentrate is subjected to the pyrometallurgical treatment, where the raw material is exposed to an intensive heating process in specially designed furnaces or kilns. Generally, the concentrate is obtained through the following stages: grinding, milling, flotation, or gravity separation. The exact pyrometallurgical process used depends on the grade of the Sb concentrate to be treated—

Table 1. Summary of the pyrometallurgical process used for Sb recovery from concentrate.

Sb Content, %	Type of Raw Material	Processing Method	Basic Chemical Reactions
5–25	sulfide	oxide volatilization (by roasting and volatilization at 1000 °C) followed by reduction smelting or reverberatory smelting of oxides	2Sb2S3 + 9O2 → 2Sb2O3 + 6SO2 2Sb + 1.5O2 → Sb2O3 Sb2O3 + 3CO → 2Sb + 3CO2 Sb2O3 + 3C → 2Sb + 3CO
25–40	sulfide	blast furnace smelting at 1300–1400 °C	Sb2S3 + 9O2 → 2Sb2O3 + 6SO2 2Sb2O3 + Sb2S3 → 6Sb + 3SO2
	sulfide	blast furnace direct reduction to metal	2Sb2S3 + 3C → 3CS2 + 4Sb Sb2S3 + 3CO → 3COS + 2Sb
45–60	sulfide	liquation at 550–600 °C, iron (Fe) precipitation alkaline smelting, or reduction smelting	Sb2S3(S) → Sb2S3(L) Sb2S3 + 3Fe → 2Sb + 3FeS 2Sb2S3 + 6Na2O + 3C → 4Sb + 6Na2S + 3CO2
>60	sulfide	iron precipitation	Sb2S3 + 3Fe → 2Sb + 3FeS
all grades	oxide	reduction smelting in reverberatory and electric or blast furnace	Sb2O3 + 3CO → 2Sb + 3CO2 CO2 + C → 2CO

[1,2,10,21,22].

Hydrometallurgical technology uses aqueous solutions and suitable chemical reactions to leach valuable metals in the pregnant leach solution (PLS). Further, the PLS is purified, and the aimed metals are recovered via different chemical and physicochemical processes. The methods for Sb bringing in the PLS and further recovery from it are similar for low-grade ores and industrial process residues.

The ten leading countries in primary Sb production in the past 10 years (2014–2023) and an estimate for 2024 are presented in Figure 1 [9,23,24,25,26,27,28,29,30,31]. According to a recent report, the USA possesses in the Idaho area 74–103 t of Sb [32]. Figure 2 offers recent data on the world reserves in the 12 richest countries with respect to primary Sb [9]. As it can be seen in Figure 1, China is the major Sb producer. Although its mine production has decreased significantly in recent years, China’s production still accounted for nearly half of the world’s Sb mine production in 2023 [9] and 2024 [31]. In addition, high-grade Sb ores are becoming scarcer, and environmental regulations are tightening, and both factors cause a further decrease in primary Sb production worldwide.

Secondary lead smelters recover secondary Sb as antimonial lead. Most of these alloys are generated by and then used again in the lead–acid battery industry [9]. However, for example, the end-of-life recycling input rate of Sb in the EU is still only 28% [8].

The anticipated increasing use of Sb and its compounds in already developed, advancing, and cutting-edge technologies indicates a growing demand for antimony. In addition, if the present extraction rate is preserved, it is expected that the extractable global resources of Sb will be exhausted by 2050 [6]. The 5-year compound annual growth rate (2020 to 2024) of the price of Sb metal increased by 37% [31].

Keeping in mind the above-mentioned points, it is no coincidence that antimony is included in the 2023 USA list of critical raw materials (CRMs) [33], as well as in the 2023 European Union list of CRMs [34]. Moreover, based on its supply risk (100% imports of primary Sb) and economic importance, Sb was included also in the previous EU CRM lists [35,36]. Australia, Canada, and Japan also consider Sb as a critical metal [21].

Because of the dissipative uses in flame retardants—one of the main end uses of Sb—and the low Sb concentration in many other products, little Sb can be recycled from these waste streams. For that reason, the industrial residues (slag, dust, anode slime, dross, etc.) generated during the extraction and recovery of Sb, Pb, Cu, gold (Au), and tin (Sn) are recognized as significant secondary resources of Sb [3,11,37,38,39].

The objective of the present review is to overview, summarize, and provide synthesized information on the industrial residues as an Sb source. The aim is to offer a structured and critical synthesis on the (i) hydrometallurgical leaching processes for antimony extraction from those residues and the effect of different process parameters (reagents’ type, solid-to-liquid ratio, temperature, and the addition of oxidizing reagents) on the effectiveness of leaching and on the (ii) leaching processes modifications and search for new leaching reagents in the past decade. Our hope is that through this text we shall assist the scientific and engineering community in (i) analytically comparing different leaching methods, identifying the knowledge gaps and needs of future research, (ii) the assessment of a viable approach to industrial residues as a valuable resource, and (iii) the selection of technological conditions for antimony extracting by leaching from residues generated by nonferrous metal mining and recovery.

2. Data Sourcing

A meticulous search using appropriate keywords (antimony, mining and metallurgical waste, recovery, extraction, leaching, acidic systems, alkaline sulfide systems, etc.) was conducted in the Science Direct database and other databases (DOAJ, PubChem, RSCI, Mendeley, BASE, etc.) to identify relevant articles, conference papers, book chapters, and published dissertations, which were then carefully reviewed. Over 500 initially identified publications were evaluated. Practically, data were used from publications in journals that are included in the most rigorous indexing databases, such as Scopus and Web of Science. The emphasis is placed on publications from the past 10–12 years, as they are considered to be built on already known knowledge. Thus, a critical summary is prepared and presented below in this article on the recent information on Sb presence in mining and metal extraction waste and means for its extraction in the PLS for further recovery as metal or Sb compounds. The extensive coverage and discussion of the recent developments in the area represents the noticeable aspect of this review.

3. Antimony Availability in Industrial Residues

The present primary production of Sb is unsustainable, outsized by growing demand from applications like lead–acid storage batteries, flame retardants, chemicals, and thermoelectric materials [40]. Therefore, recovering Sb from secondary sources is worth considering.

For example, a comprehensive worldwide study from the end of the previous century on copper refinery anode slimes pointed to the Sb content being between 0.07 and 9.4% [41]. Very useful examples of Sb content and ways of its recovery from different man-made sources are given by Anderson [37]. He pointed out the amounts of Sb contained by different materials in %: lead speiss—3.3%, lead softening skims—31.7%, copper dross obtained from flue dust treatment—9%, and Cu electrorefining slimes—8.3%.

Examples of newer data on the Sb content in different industrial residues are given in

Table 2. Examples of Sb amounts in Sb-containing copper processing residues.

Type of Man-Made Source	Sb, wt.%	As, wt.%	Bi, wt.%	Reference
Flue dust	3.1	2.8	2.8	[11]
Tin depleted anode slime	24.6	2.0	2.0	[42]
Anode slime	5.09	4.1	n.d.	[43]
Anode slime	12.62	n.d.	1.76	[44]
Flue dust	9.55	6.86	0.12	[45]
Flue dust	17.58	28.72	n.d.	[46]
Arsenic-antimony dust	20.38	53.39	n.d.	[47]
Speiss	8.8	7.7	0.003	[48]
Anode slimes from different companies worldwide	0.5–3.4	0.7–4.1	0.1–0.77	[39]
Cu mining waste—2 mines (unspecified)	0.46–1.03	0.29	n.d.	[49]
Flotation tailings from the copper-anode dressing-metallurgy	15.39	n.d.	3.9	[50]

n.d.—no data.

,

Table 3. Examples of Sb content in Sb-containing lead processing residues.

Type of Man-Made Source	Sb, wt.%	As, wt.%	Bi, wt.%	Reference
Sb dust	42.4	10.4	n.d.	[51]
Harris dross	8.2	n.d.	n.d.	[11]
Slime	63.6	4.0	3.3	[52]
Anode slime	12.62	n.d.	1.76	[44]
Matte	0.9	0.6	n.d.	[53]
Lead anode slime	22.7	23.4	0.86	[54]
Lead smelting flue dust	0.35	33.82	n.d.	[55]
Lead softening slag	17.7	4.4	n.d.	[56]

,

Table 4. Examples of Sb content in Au-Sb-rich ores.

Type of Man-Made Source	Au (g/t)	Sb, wt.%	As, wt.%	Reference
Refractory Sb-Au ore	3.6	0.3	0.4	[57]
Refractory Sb-Au ore	7.4	16.73	n.d.	[58]
Refractory Sb-Au ore	42.2	28.7	n.d.	[58]
Refractory Sb-Au ore	20.0	1.6	n.d.	[59]
Refractory Sb-Au ore	10.5	0.22	1.67	[60]
Refractory Sb-Au ore	58.8	6.30	5.50	[61]

and

Table 5. Examples of Sb content in tin recovery from industrial sources.

Type of Man-Made Source	Sb, wt.%	Bi, wt.%	As, wt.%	Reference
Tin anode slime	13.24	19.38	2.44	[62]
Slag from As–Sb dust treating plant	42.04	n.d.	n.d.	[63]
Pyrometallurgically treated tin anode slime	28.72	0.685	36.28	[64,65]

n.d.—no data.

. Data on the arsenic (As) and bismuth (Bi) content in those materials are also presented—as these elements decrease Sb leaching yield and interfere with the subsequent stages of Sb recovery from the PLS.

In addition to Sb, As, and Bi, the secondary resources presented in Table 2 contain Cu, Pb, zinc (Zn), Fe, nickel (Ni), tellurium (Te), selenium (Se), and sometimes Au, Ag, platinum (Pt), and palladium (Pd).

Antimony may be present in lead ores. It is used as an additive in the electrode material of lead–acid batteries. The electrodes of those obsolete batteries are fed in lead smelters. Due to that, Sb can be found in lead smelting waste streams—Table 3. Because of the large volume of lead production, its residues are considered a valuable source of Sb and deserve studies on their treatment for Sb recovery. In addition to Sb, As, and Bi, the secondary resources presented in Table 3 usually contain Pb, Cu, Zn, Fe, Ni, cadmium (Cd), and sometimes Ag, Au, and Sn.

Considerable amounts of Sb can be found in “refractory” gold-bearing sulfide ores. Gold cannot be leached from those ores via the standard cyanidation method. Instead, these ores are subjected to beneficiation, and antimony has to be removed from the obtained Au-Sb concentrate before Au extraction and recovery. So, such Au-Sb ores and concentrates can serve as an Sb source—Table 4.

Tailings from Sb ore processing may contain 1.21% Sb and 0.5% As, while the Sb concentrations in waste from Sb blast-furnace smelting slag are higher (0.71–3.21% Sb). The slag produced by lead addition to remove As in Sb smelting (the so-called lead removal slag) may contain Sb in the range from 5 to 45% [66]. Tailings from a stibnite flotation plant in Turkey contain 1.74% Sb and 0.32%As [67].

Araya et al. studied 16 inactive tailing deposits of the Antofagasta Region, Chile. According to their assessment, 3751 t of Sb may be recovered from those deposits [38].

Another study used data from 32% of worldwide copper refineries in order to assess the potential of refining Cu electrolytic waste to supply different metals. It is found that, in the case of 100% recovery, approximately 1497 t/year of Sb, 1632 t/yr of Bi, 4180 t/yr of selenium, and 777 t/yr of tellurium may be produced [3]. These amounts are comparable to the annual production of those metals from all primary sources [3].

All examples given above confirm that industrial residues from nonferrous metal production can be used as a source for Sb recovery. Moreover, the recovery of Sb from such residues would decrease the pollution of residence and agricultural areas that are near to mining sites. It is found that the runoff of village residential areas in an antimony mining region is polluted with different heavy metals and metalloids, and As and Sb pose the most significant health risks to the village residents [68]. Another study also pointed to soil pollution with Sb in a historical Sb mining area. The proposed Sb mobility mechanism is related to Fe reductive dissolution processes of the tripuhyite (FeSbO₄) that immobilizes Sb in soils and rocks [69].

Slag from an abandoned Sb smelter in Qinglong, Guizhou Province, China, contains 1.126% Sb that can be mobilized by rainfall and moved downstream [70].

Because of the relatively low and changeable Sb content in the above-mentioned waste materials (compared to Sb ores), generally hydrometallurgy is the suitable method to treat such sources. A general flow diagram of the Sb recovery hydrometallurgical process is presented in Figure 3. The present review summarizes and gives a general idea of the hydrometallurgical processes (with emphasis on the leaching stage), as well as their modifications proposed in the past 10–12 years, for antimony extraction from secondary non-metallic scrap resources. We would like to contribute to the selection of a suitable approach and process conditions for antimony leaching into the PLS from residues generated by the mining and recovery of nonferrous metals.

4. Antimony Leaching

As mentioned above, hydrometallurgy comprises several treatment stages. It makes use of aqueous solutions and proper lixiviants and chemical reactions to dissolve valuable metals in the pregnant leach solution. Factors that affect the leaching efficiency are generally of the lixiviant type and include the concentration, temperature, pressure, oxidation–reduction conditions, stirring conditions, and eventual pretreatment (roasting, calcination, and mechanical activation). The effects of these factors in Sb leaching are studied worldwide.

Chemical speciation studies of Sb in solids prior to leaching are vital for comprehending the form and mobility of Sb in the solid matrix. This is important for predicting the Sb leaching behavior, as well as its potential environmental impact when solid waste bears Sb. These studies include analyzing the surface and bulk chemical forms of Sb by means of X-ray absorption near edge structure (XANES) and time-of-flight secondary ion mass spectrometry (ToF-SIMS) [71]. The information gained from these analyses facilitates understanding how Sb is bound within the solid; whether it is in oxide, sulfide, or other forms; and how these forms would behave during leaching processes. The creation of antimony speciation diagrams under different conditions (using software such as HYDRA version 18 August 2009, MEDUSA version 16 December 2010, PHREEQC version 3, etc.) and pH-Eh diagrams helps in predicting the stability of Sb species and the most suitable leaching conditions [10,55,72]. Additionally, knowledge on Sb speciation is important in addressing the eventual environmental problems related to Sb leaching from contaminated sites under natural conditions [73].

Hydrometallurgical recovery of Sb from primary (mostly poor ores) and secondary sources uses mainly two lixiviant types—alkaline sulfide systems and acidic chloride systems [37].

4.1. Leaching in Acidic Systems

Singh proposed a hydrochloric (HCl) acid solution to leach Sb from Harris dross that contained 8.2% Sb at temperatures of 70–75 °C and a solid-to-liquid (S/L) ratio of 300 g/L. An antimony leaching ratio of 95.7% was achieved. Further, hydrolysis was used to obtain an Sb₂O₃ and Sb recovery rate of 73.6% [74].

Generally, it is accepted that cheaper sulfuric acid (H₂SO₄) is not a suitable leaching reagent for low-grade Sb ores and industrial process residues [37]. This is confirmed again in a recent study that compared the ability of different mineral acids to leach Sb from tailings [67]. Tailings from a stibnite flotation plant were subjected to leaching for 1 h with 4 M solutions of H₂SO₄, HCl, and nitric (HNO₃) acids at 25 °C and 90 °C. It is found that HCl possessed higher leaching efficiencies—51% and 62%. Sb was leached at the corresponding temperatures, while H₂SO₄ leached 41% and 26% of the Sb that was present in the samples, and HNO₃ leached only 14% and 18% Sb, respectively. The addition of NaNO₃ enhanced the leaching efficiency. Except for reagent type and amounts, temperature, and time, the pulp density is found as an important factor. Under optimal conditions (4.4 M HCl, 0.5 M NaNO₃, 70 °C, and 25% pulp density), in one hour, 99.88% of the Sb present in the raw material in bench-scale (1 L reactor) experiments is leached.

On the other hand, Rahimzoda and coauthors [75] claim that by leaching for 2 h with the 180 g/L H₂SO₄ solution at an S/L ratio = 1:6 and a temperature of 80 °C, 85% of the Sb that is present in the solid pyrometallurgical product obtained by 2 h of calcination with the NaCl in Sb-containing materials at 1100 °C is leached. Antimony leaching recovery from the same solid material under the same conditions was 75% and 65%, respectively, when the HCl solution and water are used. However, actually most probably in that case, “the key of success” is the introduction of chlorides in the system during the chlorination roasting process.

Sajadi and coauthors pointed out that a mixture of 5 M hydrochloric acid and 5 M sulfuric acid ensures leaching of 87% of the Sb available in the low-grade Fe₃Si₂O₅(OH)₄·Sb₃O₆(OH)·Sb₂O₃ ore, while a leaching efficiency of 85% is achieved by using a 5 M sodium hydroxide (NaOH) solution—in both cases the leaching is carried at 80 °C for 8 h [76]. The authors pointed to quicker leaching kinetics and more favorable thermodynamics in acidic conditions as key parameters. The Sb-bearing acid-based PLS ensures an opportunity for Sb recovery by electrolysis at lower temperatures and currents, compared to the alkaline PLS. However, due to the corrosion problems, the authors pointed to the alkaline process as the better one.

Acidic systems used in the industry for Sb leaching from lean ores and other Sb-bearing materials are based on HCl solutions that are often used together with ferric chloride (FeCl₃). The following generalized reactions take place (as an example for stibnite) [1,77]:

Sb₂S₃ + 6HCl → 2SbCl₃ + 3H₂S

6FeCl₃ + Sb₂S₃ → 2SbCl₃ + 6FeCl₂ + 3S^0.

Ferric chloride plays the role of an oxidizing and chloridizing agent that converts the antimony in the sulfide mineral into a chloride complex. The iron ions are written as free cations for simplicity, but in reality, in the solutions they are also existing as complex ions with chloride. Elemental sulfur is produced during that process.

The role of Fe³⁺ ion is confirmed experimentally by X-ray diffraction analysis of leach residues when the process was carried out using the H₂SO₄-NaCl-Fe(SO₄)_1.5-O₂ system [78]. The study established that the presence of ferric ions facilitates stibnite oxidation and the formation of a porous layer from sulfur.

In order to better understand the interaction of Sb-bearing material with HCl acid, Li et al. [79] studied the thermodynamics of the Sb–S–Cl–H₂O system. The species SbCl₃, SbCl₅, and Sb₄O₅Cl₂ were considered, as well as the reactions they can participate in and the corresponding stability constants. Graphs showing the distribution of different Sb species in PLS ddepending on the pH of the system (in the range from −1 to 14) and Cl^- ion concentration (in the range from 0 to 9 mol/L) were created.

However, the addition of FeCl₃ increases the leaching process costs. Excess amounts of FeCl₃ may increase Sb leaching but would interfere with the subsequent stage of metallic Sb recovery by replacement, as FeCl₃ would also oxidize metallic Fe to Fe²⁺. In the case when electrowinning is used for metallic Sb recovery, the excess FeCl₃ would decrease the Sb recovery rate on the cathode [77]. In addition, Fe³⁺ ions are good cathodic depolarizer, and their excess may facilitate the corrosion of steel equipment. The role of oxygen is to re-oxidize the ferrous ions produced [79] via the following reaction:

2Fe²⁺ + 0.5O_2(g) + 2H⁺ → 2Fe³⁺ +H₂O.

The study found that (i) in solutions without FeCl₃, the leaching mechanism is acid dissolution controlled by a chemical reaction, and (ii) in the presence of ferric ions, the Sb leaching mechanism included an acid dissolution reaction, with the diffusion of the reactant/product species occurring through the porous product layer.

When the Sb presents as an oxide in the raw material, it is directly leached with the HCl solution, and FeCl₃ is not needed:

Sb₂O₃ + 6HCl → 2SbCl₃ + 3H₂O.

Xue and coworkers proposed leaching with a 4 M HCl solution to remove antimony and arsenic from arsenic-rich copper smelter dust where Sb presented as Sb₂O₃ [46]. Under optimal conditions (90 °C, 2 h, and an L/S ratio of 6:1) 96.8% of the Sb and 97.5% of the As that are present in the dust are passed to the PLS.

Hydrochloric acid is used to leach Sb from slag generated in antimony smelting, where the Sb concentration was 4.12% and the metalloid was present in the oxide form [80]. The effects of the concentration (1, 2, 3, 4, 5, 6, 7, 8, and 9 M), solid-to-liquid ratio (from 1/5 to 1/20), temperature (room, 40, 60, and 75 °C), and time (45, 60, 90, 120, 150, and 180 min) on the extraction of Sb from the slag (<25 μm) were studied. The extraction rate of Sb exhibited a linearly increasing trend with increasing acid concentrations; however, the same trend was also observed for Fe dissolution. Based on the results, 6 and 8 M HCl solutions were chosen, and high Sb dissolution was observed, while the dissolution of all other metals together (including Fe) was lower than 30%. When the solid-to-liquid ratio decreased from 1/5 to 1/20, the extraction rate of Sb increased from 32% to approximately 45%. However, the extraction rate of Fe also increased and considerably surpassed that of Sb (Fe extraction rate > 99% at a 1/20 solid-to-liquid ratio). The extraction rate of impurities (Mn, Si, and Ca) also increased with a decreasing solid-to-liquid ratio. The antimony extraction rate increased by around 35% with time from 30 to 120 min, and then the dissolution was slowed till 180 min to reach around 80% in 3 h at 60 °C. Fe dissolution reached 90% in one hour. Antimony extraction was only around 20% at room temperature. It was increased over 2.5 times at a temperature of 60 °C. Under the optimum leaching conditions (8 M HCl, 75 °C, 3 h, and intensive stirring), 91.19% of the Sb available in the slag was leached. Based on shrinking core modeling and the calculated activation energy (46.75 kJ/mol), it is concluded that leaching is a chemically controlled process. The obtained PLS is suitable to produce clean Sb-bearing precipitates (consisting mainly of Sb₂O₃) by hydrolysis achieved by the NH₄OH addition.

The use of a chlorine/SbCl₅-containing solution from Sb recovery via electrowinning in a chlorination–oxidation treatment facilitates Sb leaching and obtaining the SbCl₃ solution from the complex raw material [81]. The process achieved 99.5% leaching of the Sb available but at the expense of the application of a highly corrosive leaching reagent. However, the authors do not comment on the possible unavoidable problems associated with working with chlorine, which is hazardous to the environment and human health. Although the Cl₂/SbCl₅ anolyte obtained after Sb metal electrowinning is recycled as an oxidant in the leaching process, the corrosiveness of the electrolyte is not discussed. The described drawbacks of the chlorination leaching process may hinder this technology’s industrial acceptability.

Tian and coauthors proposed the use of ozone as an oxidizer in the HCl system with the aim to facilitate Sb leaching from complex sulfidic antimony raw materials and achieve maximum Sb dissolution at minimum dissolution for iron [82]. The main reactions proceeding can be expressed in the following way:

Sb₂S₃ + 2iCl⁻ + 6H⁺ + 3O₃ → 3S⁰ + 2SbCl^3-i + 3H₂O + 3O₂

FeS₂ + 2H⁺ + O₃ → Fe²⁺ + 2S⁰ + O₂ + H₂O

Sb₂S₃ + 2iCl⁻ + 3H⁺ + 12O₃ → 3HSO₄⁻ + 2SbCl^3-i + 12O₂

FeS₂ + H₂O + 7O₃ → Fe²⁺ + 2HSO₄⁻ + 7O₂.

It can be seen that S²⁻ in stibnite is oxidized to elemental sulfur. That is why sulfur remains in the residue, while antimony dissolves in the form of Sb³⁺. The scientists found that the antimony leaching efficiency increased with increasing HCl concentration (from 3.0 mol/L to 4.5 mol/L) and temperature (from 35 °C to 65 °C). Under these conditions the rate of iron dissolution did not increase. That is why in 4 h, Sb extraction of 94.3% is achieved with only 2.3% dissolution of Fe under the following optimal conditions: 4.5 mol/L hydrochloric acid, an ozone flow rate of 2.0 L/min, an L/S ratio of 8:1 and a temperature of 65 °C. Further, the researchers found that the shrinking core model described the leaching process [83]. The diffusion-controlled model (with an apparent activation energy of 6.91 kJ/mol) fits Sb leaching at low temperatures (15–45 °C). At high temperatures (45–85 °C) the mixed-controlled model (with the apparent activation energy of 17.93 kJ/mol) better reflects the process. The same working group proved that the use of ozone in Sb leaching from antimony-bearing refractory gold concentrates significantly improved Sb leaching from the material, as well as Sb and Au separation. Gold is concentrated and enriched in the leaching residue [84]. It is found that the Sb leaching efficiency increased with increasing HCl concentration, temperature, L/S ratio, and stirring speed. Under optimal conditions (HCl 3.0 mol/L, L/S—10:1, 85 °C, and active stirring) 93.75% of the Sb available in the material is extracted, and the residue is enriched in Au to reach 61 g/t at an initial concentration of 55 g/t.

In order to avoid the use of ozone, which is very corrosive and difficult to work with, NaNO₃ is proposed as an oxidizer that facilitates the dissolution of metals from the sulfide minerals in the HCl solution. Semi-pilot tests in a 10 L reactor (under the following conditions—4.4 M HCl, 0.5 M NaNO₃, 70 ◦C, and 25% pulp density) achieved leaching efficiencies of 98.30% for Sb, 98.00% for Fe, and 92.00% for As from stibnite flotation tailings [67].

Studies with the proposed reagents and oxidants pave a way for optimizing Sb leaching from a complex and with a low content of man-made Sb sources.

Table 6. Summary of some experimental hydrometallurgical studies on acidic antimony leaching.

Raw Material; Sb Content, %	Operating Conditions	Sb Recovery in PLS, %	Reference
Cu smelter dust, 17.58%	4 M HCl, L:S = 6:1, 90 °C, and 2 h	97.53	[46]
Stibnite flotation tailings, 1.74%	4.4 M HCl, 0.5 M NaNO3, pulp density 25%, 70 °C, and 1 h	99.88	[67]
Low-grade ore, n.a. % Fe3Si2O5(OH)4·Sb3O6(OH)	5 M HCl, 80 °C, and 8 h	87	[76]
Slag from Sb smelting, 4.12%	8 M HCl, 75 °C, and 3 h	91.9	[80]
Stibnite concentrate, 61.85%	4 M HCl, L:S = 5:1, 85 °C, and 2 h (Cl2/SbCl5 solution)	99.5	[81]
Complex sulfidic Sb ore, 58.57%	4.5 M HCl, L:S = 8:1, 65 °C, and 2 h (O3 2 L/min)	94.3	[82]

summarizes some experimental hydrometallurgical studies on acidic antimony leaching.

It could be briefed that the use of the HCl solution with a concentration of over 4 M, a pulp density in the range of 25–35%, a moderate temperature increase of up to 60–90 °C, stirring, the presence of an oxidizer (especially for sulfide raw materials), and a leaching time of 2–4 h, in general, may be considered as suitable conditions for antimony leaching. Easier downstream separation of antimony by precipitation (e.g., Sb₂O₃ via hydrolysis) can be pointed out as an advantage of the acid leaching system.

However, due to the fact that HCl is a highly corrosive reagent, it requires special equipment—a prerequisite for higher costs. In addition, acid use is chemically more complex [10]. For these reasons, the alkaline–sulfide system is employed more often industrially in Sb leaching.

4.2. Leaching in Alkaline Sulfide Systems

In the alkaline leaching system, the lixiviant is a mixture of sodium sulfide (Na₂S) and sodium hydroxide (NaOH). When stibnite contacts with this solution, sodium thioantimonite (Na₃SbS₃) is formed as described by the following reactions [37]:

Na₂S + Sb₂S₃ → 2NaSbS₂

NaSbS₂ + Na₂S → Na₃SbS₃.

The dissolution of elemental sulfur (naturally available in minerals) in NaOH forms sulfur species that can also serve as a lixiviant for antimony in alkaline sulfide leaching. Sulfide (S²⁻), sodium thiosulfate (Na₂S₂O₃), and sodium polysulfides (Na₂S_x) are created, as presented by the following generalized reactions:

4S⁰ + 6NaOH → 2Na₂S + Na₂S₂O₃ + 3H₂O

(x − 1)S⁰ + Na₂S → Na₂S_x (where x = 2 to 5).

Polysulfide oxidizes sodium thioantimonite, leading to the formation of sodium thioantimonate (Na₃SbS₄) that is the major species found usually in the solution:

Na₂S_x + (x − 1)Na₃SbS₃ → (x − 1)Na₃SbS₄ + Na₂S

Generally, NaOH is added to prevent Na₂S hydrolysis by the following reactions [2]:

Na₂S + H₂O → NaHS + NaOH

NaHS + H₂O → H₂S + NaOH.

The overall reaction is Na₂S + 2H₂O → H₂S + 2NaOH.

When there is insufficient Na₂S in the leach solution, NaOH also solubilizes stibnite by producing alkaline thioantimonides and oxothioantimonites [2]:

Sb₂S₃ + 2NaOH → NaSbS₂ + NaSbOS + H₂O.

It has to be mentioned that alkaline sulfide leaching is very selective for antimony, gold, arsenic, mercury, and tin minerals. However, it still exhibits some drawbacks. For instance, due to the fact that stibnite occurs with other sulfide minerals (galena, sphalerite, pyrite, etc.), NaOH leaches these minerals to form correspondingly soluble plumbite (Na₂PbO₂), zincate (Na₂ZnO₂), and iron hydroxide (Fe(OH)₂) that can cause a certain loss of the reagent [10]. In addition, alkaline sulfide leaching creates by-products such as Na₂S₂O₃ and Na₂SO₄ that may pose disposal problems [77]. Sodium sulfide is classified as harmful if swallowed; toxic upon contact with the skin, causing severe skin burns and eye damage; and toxic to aquatic life [85]. Toxic H₂S may be evolved when working with Na₂S. These safety issues require more work safety and environment protection measures when it is used. For these reasons efforts are being made for alkaline sulfide leaching modification and optimization.

By using the response surface methodology–central composite face-centered design, Awe and coauthors found that the leaching process of a complex copper concentrate bearing 1.69% Sb strongly depends on the sulfide ion concentration and reaction temperature, while the dependence on solid concentration is not significant [86].

Alkaline sulfide leaching is proposed for Sb recovering from lead silicate slag containing 6.5% Sb in addition to 19.4% Pb, 10.7% Cu, and 3.4% As. An increase in the Sb leaching rate is observed with increasing NaOH concentration from 10 g/L to 30 g/L and temperature from 70 to 100 °C. It is found that 83% of the Sb that is present in the material can be extracted for 24 h at 100 °C by using a lixiviant containing 30 g/L NaOH and 30 g/L S²⁻ ions [48].

Leaching with Na₂S is proposed for extracting Sb from tin anode slime obtained by soda roasting–alkaline leaching and containing, in wt.%, Sb—13.24, Bi—19.38, and As—2.44 [64]. The proceeding chemical reaction causing Sb dissolution produces liquid sodium thioantimonate and can be expressed as follows:

NaSb(OH)₆ + Na₂S → 6NaOH + Na₃SbS₄

It is found that the leaching efficiency is significantly affected by the leaching temperature and the Na₂S concentration, while the leaching time and the L/S ratio are not as influential. Under the optimum conditions of 0.7 mol/L Na₂S, 85 °C, and L/S = 14:1, 98% of Sb was extracted from the Sn anode slime in 120 min. The evaporation–crystallization of the produced liquid sodium thioantimonate converted it into Na₃SbS₄(H₂O)₉ crystals, and the recovery rate of this process was 93.83%.

A microwave-assisted heating alkaline leaching process for Sb (that is presents as tetrahedrite) from Sb-bearing copper concentrates is proposed [87]. It is found that antimony dissolution is very selective and can be described by the following reaction:

Cu₁₂Sb₄S_13(s)+2S²⁻ → 5Cu₂S_(s) + 2CuS_(s) + 2Sb₂S₄²⁻

As can be seen, the copper present in tetrahedrite is not leached. It is transformed into a mixture of solid copper sulfides which facilitates Cu recovery. It is found that microwave-assisted heating enhanced leaching while decreasing the needed temperature and leaching time. Under the optimal conditions (leaching solution containing 250 g/L Na₂S and 60 g/L NaOH and 140 °C achieved with microwave-assisted heating in 2 h) the leaching efficiency of Sb was approximately 96%. While decreasing the leaching temperature and time means decreased energy consumption, the question about the energy needed for using microwave radiation as heating technology is risen. Another question is related to the technology’s scalability to industrial conditions, since in the experiments described small amounts of the sample (2.5 g) and lixiviant (50 mL) are used. The positive result of this study is suppressed copper leaching—the Cu concentration in the PLS was lower than 2 mg/L. Copper concentrate suitable for metallurgical processing is obtained since its Sb content was decreased from 1.1 wt% to <0.2 wt%. This result provokes the interest of the scientific and industrial community and will certainly lead to further research in this direction.

The leaching of antimony with the Na₂S solution is proposed as a first step in a process aimed at the recovery of gold from the refractory gold ore. The leaching efficiency for antimony, gold, and arsenic was 96.64%, 1.44%, and 0.41%, respectively, under the optimum conditions for antimony dissolution (Na₂S 50 g/L, NaOH 20 g/L, 50 °C, and a liquid-to-solid ratio of 1.5 L/kg) [61].

Sulfide–alkaline leaching is suggested as a first stage of a metallurgical process aimed to recover Sb and Au from gold–antimony concentrate containing high amounts of antimony and arsenic [88]. In order to decide on proper leaching conditions, the authors constructed an Eh-pH diagram that shows different stable species of the Sb–Au–S–H₂O system. It is found that in this case, besides the Na₂S concentration, the L/S ratio has a significant influence on the leaching process. At the optimal leaching conditions (a Na₂S concentration of 61 g/L, a NaOH concentration of 16.5 g/L, pH > 13, an L/S ratio of 4.5:1, 50 °C, and 3 h), the Sb extraction of 99% is achieved.

After studying the use of acidic lixiviants for Sb leaching from stibnite flotation tailings [67], Dembele and coauthors tested alkaline sulfide-based lixiviants [89]. The effects of the alkalizing reagent (NaOH and KOH) and its concentration (from 0.5 M to 6 M), pulp density (20%; 25% and 30%), temperatures (25, 50, 70 and 90 °C), time, and stirring speeds (300 rpm and 400 rpm) on Sb and As leaching are studied. For both alkalizers increasing the concentrations caused an increase in Sb leaching efficiency. Better results are obtained with NaOH. Increasing the lixiviant concentration increased As leaching by up to 2 M NaOH and KOH. It is found that Sb leaching efficiency decreased slightly after 1 h, while As leaching efficiency decreased significantly after 2 h. Temperature significantly increases Sb leaching from 25 °C to 50 °C, within 1 h, and then a slight increase is observed beyond this range. The optimum Sb leaching efficiencies (in 1 h) were 71.54%, 93.20%, 96.53%, and 97.14% at 25 °C, 50 °C, 70 °C, and 90 °C, respectively. Better results with respect to Sb leaching are obtained at a lower stirring rate.

The authors conducted process optimization using the Design Expert program, entering the appropriate data for the variable parameters (Na₂S and NaOH concentrations and pulp density). The program suggests different solutions, and the optimum conditions chosen by the authors for the validation tests were 25% pulp density, 0.97 M Na₂S, and 2.5 M NaOH with 95.31% Sb recovery and the value of desirability of 0.942. Duplicate trials were run at 90 °C and 400 rpm for 1 h to compare the predicted and experimentally obtained Sb recovery results. The test result (97.36% Sb) was slightly higher (by 2%) than the predicted value. Bench-scale (1 L reactor) and semi-pilot-scale (10 L reactor) results confirmed the optimum chosen process parameters. Under the optimum parameters (25% pulp density, 0.97 M Na₂S, 2.5 M NaOH, 70 °C, 300 rpm, and 1 h) the antimony dissolution rates were 99.13% and 97.00% at the bench and semi-pilot scales, respectively.

Table 7. Summary of some experimental hydrometallurgical studies on alkaline antimony leaching.

Raw Material; Sb Content, %	Operating Conditions	Sb Recovery in PLS, %	Reference
Lead silicate slag, 6.5%	0.75 M NaOH, 100 °C, and 24 h	83	[48]
Refractory Au ore, 6.30%	0.5 M NaOH, 1 M Na2S, L:S = 1.5:1, 50 °C, and 1.5 h	96.64	[61]
Stibnite flotation tailings, 1.74%	2.5 M NaOH, 0.97 M Na2S, pulp density = 25%, 70 °C, and 1 h	Bench scale 99.13 Semi pilot scale 97.00	[67]
Tin anode slime obtained by soda roasting, 13.21%	0.7 M Na2S, L:S = 14:1, 85 °C, and 2 h	98	[62]
Low-grade ore, n.a. % Fe3Si2O5(OH)4·Sb3O6(OH)	5 M NaOH, 0.5 M Na2S, 80 °C, and 8 h	85	[76]
Complex Cu concentrate, 1.69%	0.2 M NaOH, 80 °C, and 20 h	52	[86]
Sb-bearing Cu concentrate, 1.04%	1.5 M NaOH, 4.5 M Na2S, 140 °C, and 2 h (microwave heating)	96	[87]
Au-Sb concentrate, 19.18%	0.4 M NaOH, 1.1 M Na2S, L:S = 4.5:1, 50 °C, and 3 h	99	[88]

outlines some experimental hydrometallurgical studies on alkaline antimony leaching.

Because of the different characteristics of the Sb-bearing materials, the span of the (found by the different studies) optimal conditions for Sb leaching in the alkaline leaching system is relatively large. It may be loosely summarized that the NaOH concentration in the range of 0.5–2.5 mol/L, the sulfide concentration in the range of 1–2 mol/L, and the temperature in the range of 70–100 °C would ensure suitable conditions to extract in 2–3 h at least around 90% of the Sb that is present in the raw material.

Based on the studies presented in this paper, acidic and alkaline antimony leaching processes may be compared, as shown in

Table 8. Comparison of acidic and alkaline antimony leaching.

Leaching System; Parameter	HCl-Based	NaOH + Na2S Based	Comments
Temperature	65–90	70–100	Keeping in mind both temperature and process duration, it could be expected that there would not be a big difference in energy consumption.
Time	2–3	1.5–2
Reagent’s consumption	>4 M	0.5–2.5 M NaOH 1–2 M Na2S	Keeping in mind both reagents’ consumption and price, it could be expected that there would not be a big difference in the costs related to the reagents.
Reagent’s price, USD/kg	0.2–0.30 1	NaOH 0.43–0.70 2 Na2S 0.41–0.43 3
Sb leaching efficiency	87–99%	85–99%	Generally, the addition of an oxidizing reagent is needed to achieve efficiency > 90% by HCl-based systems.
Safety issues and hazards according to the GHS 4	H290 H314 H335	NaOH: H290; H302; H314 Na2S: H302; H311; H314; H400	Proper measures for protecting the workers’ health and the environment have to be considered with both leaching systems.
Selectivity with respect to Sb	Low	High	A decrease in the number of procedures and costs for further Sb recovery from the PLS from alkaline solutions is expected.
Sb separation from the PLS	Hydrolysis, conversion, and electrowinning	Precipitation, crystallization, and electrowinning
Equipment corrosion	High	Low	Although the H290 code is given for both systems, it is well known from corrosion science that the acidic medium is considerably more corrosive compared to the alkaline one for general-purpose leaching equipment, especially in the pH ranges used in Sb leaching.
Scalability	Difficult, mainly due to corrosion problems	Relatively easy	Alkaline-based leaching predominates in real industrial applications.

¹ https://www.made-in-china.com/products-search/hot-china-products/Hydrochloric_Acid_Price.html; (accessed on 25 May 2025) ² https://www.alibaba.com/showroom/sodium-hydroxide-price.html; (accessed on 25 May 2025) ³ https://www.made-in-china.com/products-search/hot-china-products/Sodium_Sulfide_Price.html; (accessed on 25 May 2025) ⁴ United Nations, Globally Harmonized System of Classification and Labelling of Chemicals (GHS), 2021; H290—may be corrosive to metal; H302—harmful if swallowed; H311—toxic in contact with skin; H314—causes severe skin burns and eye damage; H335—may cause respiratory irritation; H400—toxic to aquatic life.

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4.3. Leaching in Other Solutions

Leaching with the FeCl₃ solution is proposed as an environmentally friendly approach for stibnite treatment [77]. Antimony is leached, producing SbCl₃, and S⁰ and Fe³⁺ ions are reduced to Fe²⁺ ions. The leaching of Sb increased with the increase in the applied amount of FeCl₃ until the optimum is achieved. It is found that the excess amounts of FeCl₃ interfere with the subsequent stage of solid Sb recovery by replacement with iron. Under the optimum conditions (L/S = 8:1, 50 °C, the excess stoichiometric coefficient of FeCl₃ to Sb₂S₃ equal to 1.2, and 1 h) the Sb leaching efficiency was 98.53%.

Mixtures of Lewis or Bronsted acids and bases that give rise to liquids are named deep eutectic solvents (DESs). DESs can be classified into four categories [90]:

“Type I—Quaternary ammonium salt and metal chloride;

Type II—Quaternary ammonium salt and metal chloride hydrate;

Type III—Quaternary ammonium salt and hydrogen bond donor (HBD);

Type IV—metal chloride hydrate and HBD”.

DESs exhibit low volatility and consequently low vapor pressure and high thermal stability. DES toxicity is low, compared to that of volatile organic compounds, and some DESs are biodegradable. They are cost-effective alternatives of ionic liquids. DESs consist of large asymmetric ions with low lattice energy. They are prepared mainly by combining quaternary ammonium salt with an HBD—usually by mixing the components with moderate heating.

Type III DESs prepared by blending choline chloride and different HBDs (ethylene glycol, malonic acid, and thiourea) have been studied for their ability under different conditions (a temperature of 60–100 °C, leaching time—5 min to 24 h, and the percentual fraction of the solid component: 12.18–30.45) to leach Sb from materials taken from the Čučma tailings pond, Rožŝava district, Slovakia [91]. The material contained less than 0.7% antimony. It is found that the mixtures of choline chloride, malonic acid, choline chloride, and thiourea do not ensure high leaching efficiency under all studied conditions. Leaching with malonic acid as an HBD achieved 25.05% at 100 °C for 24 h. Leaching with thiourea as an HBD achieved a maximum of 25.97% at 80 °C for 24 h. The mixture of choline chloride and ethylene glycol ensured practically 100% Sb leaching for 4 h at 100 °C, a solid content of 30.45%, and the addition of 3 g of iodine as an oxidant, thus highlighting the DESs as a promising lixiviant for Sb sustainable recovery from mining tailings and other complex secondary raw materials. The use of DESs minimizes the use of harsh chemicals and reduces the environmental footprint. In addition, DES extraction may be combined with the electro-deposition of Sb directly from the PLS, thus rationalizing the process of metal recovery. For instance, a process that simultaneously dissolves and electro-deposits using a mixture of choline chloride and ethylene glycol is proposed for retrieving lead from Pb-based hybrid organic–inorganic perovskite photovoltaic devices [92].

However, DESs often suffer from high viscosity, especially when the water content is low. Mass transport is hampered by high viscosity, and the occurring chemical reactions are slowed down. In the further stage of Sb recovery from the PLS, high viscosity makes solid–liquid separations more difficult.

Future studies have to be aimed at finding DES compositions that would ensure the maximum amount of Sb leaching at lower temperatures, thus decreasing the carbon footprint. The kinetics of the leaching process needs further study with the aim to decrease the leaching time. In addition, the scalability and economic feasibility of the extraction process need further assessments.

Solvometallurgy is the term used to designate the extraction of metals from raw materials (ores, industrial residues, production scrap, etc.) via the use of non-aqueous, anhydrous solutions instead of water-based solutions. The lixiviants can be acidic extractants, acid-saturated neutral or basic extractants, chelating extractants diluted in nonpolar organic solvents, halogens in organic solvents, or acid-saturated polar solvents. Solvometallurgy involves low-temperature processes that saves energy. It often uses green environmentally friendly solvents. Solvent leaching is frequently more selective when compared to acidic leaching, leading to reduced acid consumption and fewer purification steps [93]. Solvometallurgy possesses several advantages: (a) very limited consumption of water and very low volumes of generated wastewater, (b) reduced consumption of acids, and (c) reduced energy consumption. (d) No silica gel is formed, so solvometallurgy is suitable to treat ores rich in soluble silica. (e) The leaching and solvent extraction process can be combined in a single step, leading to more simplified process flow sheets and process intensification. (f) Solvent leaching is frequently more selective than leaching with acidic aqueous solutions, and the PLS needs fewer purification steps. (g) The method is suitable for the treatment of low-grade ores, mine tailings, and industrial process residues [93].

Solvent leaching is studied as a means to leach antimony from the lead-rich dross of a lead smelter with the aim to recover antimony as an antimony oxide chloride (Sb₄O₅Cl₂)—a material that can be used to prepare anodes in aqueous chloride batteries or as a component in flame retardants [94]. The dross contained 30 wt.% Pb, nearly 30% Sb, and relatively low amounts of zinc, iron, and tin. Different lixiviants were studied—hydrochloric acid dissolved in organic solvents such as ethanol, ethylene glycol, 1-octanol, and Aliquat 336 chloride. All the studied solutions leached similar amounts of antimony (60–76%). However, the lowest dissolution of lead (∼0.1%) was attained by hydrochloric acid dissolved in ethanol or 1-octanol. Ethanol is considered an environmentally friendly solvent, and it is much cheaper than 1-octanol. A lower concentration of hydrochloric acid in ethanol is needed in comparison with the HCl concentration in water to achieve high leaching yields of antimony. At the same time the amount of co-leached Pb is considerably lower compared to the case when HCl is dissolved in water. Batch experiments (with a 1 L leaching reactor at 25 °C for 6 h) achieved an antimony leaching efficiency of 90%, while the lead leaching efficiency was 0.4%. The PLS contained 28 g/L Sb and only 100 mg/L Pb. The dissolved Sb from the PLS was precipitated as high-purity Sb₄O₅Cl₂ by hydrolysis accomplished by water addition. The ethanol remained in PLS and was recovered by distillation. The high Sb concentration in the PLS allows for direct Sb electrowinning. The use of ethanol and the valorization of an industrial process residue add to the environmentally friendliness and sustainability of the process.

Although solvometallurgy is a promising approach for metals leaching, several problems have to be solved before the application of these processes on a large scale in industry. Organic extractants and solvents are more expensive compared to water and mineral acids. Technologies are needed for the recycling of these solvents and extractants at the end of the process. Careful selection of solvents is required—they must have low flammability, low toxicity, and a low environmental impact. The solvents used have to be biodegradable. At the same time, they have to ensure high Sb leaching efficiency.

The number of works describing solvometallurgical processes is very low, compared to the large number on hydrometallurgical and pyrometallurgical processes. Despite its attractiveness, the process of solvent Sb leaching for the present is studied only at the laboratory scale. Its scalability has to be investigated. Up to now, no solvometallurgical processes applied at the industrial scale have been revealed [93]. The Technology Readiness Level (TRL) of solvometallurgical processes is still low. This is a disadvantage for short-term implementation, including in Sb leaching, but provides an opportunity for further research, development, and innovation.

4.4. Bioleaching

Bio-hydrometallurgy employs microorganisms to extract metals from ores, concentrates, or metal-bearing wastes. Due to its moderate capital and low operating costs, bioleaching is considered a cost-effective alternative for metal production, especially from low-grade ores, tailings, and residues of mining and metallurgical plants. The commercial use of bio-oxidation and bioleaching includes the extraction of copper, nickel, gold, zinc, and cobalt. Bioleaching that applies iron- and/or sulfur-oxidizing bacteria is an established process for sulfide ores, which hints to the idea that stibnite and other Sb sulfides may be leached with the aid of microorganisms. However, bioleaching has not yet been used at the industrial level for Sb recovery [2].

Nevertheless, studies have been conducted at the laboratory level. In general, Acidithiobacillus ferrooxidans, as well as Leptospirillum ferrooxidans and Acidithiobacillus thiooxidans, is being studied for its ability to leach stibnite ores [95,96,97].

Acidithiobacillus ferrooxidans oxidizes stibnite to antimony (III) sulfate, which further easily hydrolyses to form an insoluble antimony (III) oxosulfate, as shown below [98]:

Sb₂S₃ + 6O₂ + 2H₂O → (SbO)₂SO₄ + 2H₂SO₄

Hafeez et al. [95], using the same bacterial genus, leached 73% of the Sb that is present in stibnite in 120 days. Loni et al. [97] found that the bacterial strain Paraccocus versutus XT0.6 that was isolated from the Xikuangshan antimony mine is capable of dissolving Sb³⁺ in minerals/rocks by changing the pH. The strain XT0.6 can oxidize dissolved Sb³⁺ aerobically and anaerobically.

Čerňanský and coauthors studied the capability of the fungal strain Aspergillus niger to leach Sb and As from mining waste taken from the Slovinky tailing impoundment (Slovakia) [99]. The concentration of Sb in the samples was in the range of 225.3–285.7 mg/kg. It is found that Sb amounts in the range from 10.8 to 13.7% of the initially available in the raw material were leached in 21 days of cultivation.

Another study pointed to the opportunity to extract additional amounts of Sb left in the residue after Na₂S-NaOH leaching of tetrahedrite concentrate. The maximum leaching efficiency of 92.8% with respect to residual Sb is achieved by using moderately thermophilic cultures (Acidithiobacillus caldus, Sulfobacillus thermosulfidooxidans, Leptospirillum ferriphilum, and Ferroplasma spp.) The total Sb extraction of 99.4% is achieved when alkaline and bacterial leaching are consecutively applied to the tetrahedrite concentrate [100].

Bio-oxidation and bioleaching have also been used as a pre-treatment process for Sb removal from sulfide-containing gold or silver ores. Tsaplina et al. [101] studied the ability of a consortium of thermo-acidophilic microorganisms (belonging to the genus Sulfobacillus, Leptospirillum, and Ferroplasma) to leach Sb from different Au-Sb ores (containing 0.84 to 29.95% Sb sulfide). It is found that independent of the antimony concentration in the solid samples, after adaptation to a specific ore, the microorganisms grew actively at 39 °C and oxidized and leached the studied gold–antimony ores. Up to 86% of the Sb may be leached prior to the gold cyanidation process.

In order to improve the efficiency and stability of the bio-oxidation process in Sb-bearing gold sulfide ores (containing 4.3% stibnite), Muravyov [96] proposed a two-stage bio-oxidation process with preliminary ferric leaching of the concentrate for 5 h at 80 °C and a pH range of 1.15–1.35. The first bioleaching stage is carried out at 50 °C and the second at 39 °C. In the bioleaching step a moderately thermophilic culture of acidophilic microorganisms, mainly of the genus Sulfobacillus, was used. The residence time of the suspension in each bioreactor was 2 days. The residue from bio-oxidation leaching was submitted for gold cyanidation with 82% Au dissolution.

De Carvalho et al. [102] studied the ability of Acidithiobacillus ferrooxidans to oxidize gold-bearing sulfide concentrate (3.16% Sb) and a flotation tailing as a pre-treatment step prior to gold cyanidation. At optimum conditions of bio-oxidation (pH 1.75, 10% pulp density, a temperature of 32 °C, a stirring speed of 150 rpm, 5.0 g/L Fe²⁺, and within 40 days), gold leaching was 85% for both samples. On the contrary to the previous investigations, this study disclosed that the dissolution of Sb was low due to its co-precipitation with iron. Antimony is found mainly in the solid product of the bioleaching process as inert sulfide particles. Therefore, the authors concluded that bioleaching is not a viable option for the removal of Sb from gold ores.

Another recent study also pointed to the relatively low leaching rate of Sb from a polymetallic Ag-Cu-Sb-siderite ore [103]. The ore was ground to a grain size of <100 μm, upgraded by froth flotation and acid leaching to produce a siderite-free tetrahedrite concentrate. Bioleaching of the concentrate was carried out in a sulfate medium (with Ph ~ 2) by iron-oxidizing bacteria, namely Acidithiobacillus ferrooxidans, Acidithiobacillus ferrivorans SS3, and Leptospirillum ferriphilum, under aerobic conditions at 25 °C for 210 days in order to monitor the concentration changes in individual metals and metalloids in the leachate. It is found that about a 12.7% Sb leaching yield is achieved on the 30th day of bioleaching, and subsequently the process efficiency decreased.

The information on the bioleaching of stibnite is summarized in

Table 9. Some examples of bioleaching of stibnite.

Microorganism	Leaching Conditions	Sb Recovery in the PLS	Reference
At. ferrooxidans	pH 2, 37 °C, and 120 days	73%	[95]
Sulfobacillus	pH 1.6–2.0, 50 °C and 39 °C, and 4 days	n.d.	[96]
Aspergillus niger	pH 8.49, 21 days	10.8 to 13.7%	[99]
Mesophilic: Acidithiobacillus ferrooxidans, Acidithiobacillus thiooxidans, and Leptospirillum ferrooxidans Moderately thermophilic prokaryotes: Acidithiobacillus caldus, Leptospirillum ferriphilum, and Sulfobacillus	initial pH 1.8, 34 °C for mesophiles and 45 °C for thermophiles, and 17 days	6.6%	[100]
Sulfobacillus, Leptospirillum, and Ferroplasma	pH 1.8–1.9, 39 °C, and 14 days	20.6–86.2%	[101]
Acidithiobacillus ferrooxidans	pH 1.75, 32 °C, and 40 days	<13%	[102]
At. ferrooxidans, At. ferrivorans SS3, and Leptospirillum ferriphilum	pH ≅ 2, 25 °C, and 30 days	12.7%	[103]

n.d.—no data.

.

Although it is very attractive from the environmental protection point of view, bioleaching of antimony-containing materials has not yet found its application in real industrial conditions. The main reasons for this are the low degree of extraction of antimony in production solutions, compared to conventional technologies, and the need for a long period of contact of the material with microorganisms. In addition, the implementation in real practice is hampered by the need to strictly maintain the temperature within a narrow temperature range but elevated compared to ambient temperature. An additional obstacle is the need to isolate and cultivate suitable microorganisms and create conditions to maintain their vital activity. Further research is required to overcome the mentioned obstacles.

Based on the above-described works, a general decision framework for leaching route selection is proposed in Figure 4.

In order to recover Sb from the PLS, selective precipitation/cementation, solvent extraction/re-extraction, ion exchange/elution, and electrowinning are applied. These stages of hydrometallurgical recovery of Sb are out of this paper’s scope and deserve to be a subject for another work.

5. Conclusions and Outlook

Antimony-containing residues from nonferrous metal extraction can serve as a significant secondary source for Sb recovery that will help in mitigating the expected worldwide shortage of Sb in the future. In addition, the health and environmental risks posed by such waste materials will be alleviated. Further studies on Sb recovery from such sources and development of technological processing flow sheets are highly needed. Except for technological constrains, Sb dissipative use and awareness of Sb-bearing waste as a source are the other general obstacles in using residues of nonferrous metal production.

Applying the pyrometallurgical methods to treat such man-made sources exhibits shortcomings, like the high energy consumption and eventual environmental pollution with CO₂, SO₂, antimony, and arsenic compounds.

Hydrometallurgical methods are more suitable for recovering Sb from waste generated by nonferrous metal extraction due to the following advantages: (i) possibility to find selective leaching conditions with respect to Sb dissolution, (ii) efficiency when applied to low-grade Sb-bearing sources, and (iii) adaptability and scalability, since hydrometallurgical processes can be easily adapted to different volumes and compositions of materials to be treated. The lixiviants, temperature, pH, oxidation–reduction conditions, solid-to-liquid ratios, etc., can be tuned to optimize Sb recovery from different sources. (iv) In general, hydrometallurgical methods are considered more environmentally friendly compared to pyrometallurgical ones due to the lower energy needed (operation at lower temperatures), lower gas emissions, a lower carbon footprint, and the availability of means to stabilize the waste streams when formed. (v) This technology route gives a wide opportunity to search for and test environmentally benign lixiviants.

Studies on Sb leaching from nonferrous metal extraction waste have been carried out mainly using HCl acid-based media and alkaline—Na₂S-NaOH-based media. In the majority of the cases, around 90% of the Sb present in secondary resources can be leached using hydrochloric acid or alkaline sulfide-based lixiviants.

Leaching in an acidic solution can be carried out at a lower temperature compared to alkaline leaching but generally needs a longer time; thus it can be considered that both technologies would require similar amounts of energy. Hydrochloric acid is cheaper than the reagents needed for alkaline leaching, but its higher concentration in the lixiviants is required. Furthermore, in order to achieve higher Sb leaching yields, the addition of oxidizing reagents is needed for sulfidic raw materials. This practically equalizes the costs for reagents for both leaching methods. Alkaline leaching offers more selective dissolution of antimony from complex waste matrices.

Recent advances in Sb leaching include the use of different oxidants to facilitate the process, the optimization of leaching conditions, studies on the leaching kinetics, and the application of microwave-assisted heating to limit the formation of undesired by-products.

The choice of the leaching process depends on the type and composition of the Sb-bearing raw material, the presence of other metals, and the subsequent planned treatment for Sb extraction from PLS. Leaching in acidic systems is recommended when the raw material contains Sb mainly as oxides. A lower corrosion risk for the equipment and easier scalability have made alkaline leaching the predominating method in real industrial applications.

Problems with the handling of impurities in a more efficient manner and reagent recycling still depend on their solution for both leaching systems and need further studies.

The use of eutectic solvents and ionic liquids is able to offer lower volatility, reusability, high selectivity, and the exploration of bacterial or fungal cultures that either directly solubilize antimony or create acidic/oxidizing conditions for Sb leaching and can be considered as new frontiers in the hydrometallurgical extraction of Sb from man-made sources. These routes exhibit a potential for lower energy consumption and reduced use of harsh chemicals and should be further insistently explored.

The high viscosity of DES hinders mass transport and thus slows reactions, so solving this problem, together with the process’s scalability and economic feasibility, can be pointed as future priority research directions related to DES technology.

Significantly increasing the Sb leaching yield in the PLS, isolating/determining more effective microorganisms, maintaining their vitality, and decreasing the leaching time and cultivating temperature can be pointed as the main problems that have to be solved before bioleaching implementation at the industrial scale.

The application of solvometallurgy in Sb extraction from raw materials is a promising technology. However, the Technology Readiness Level (TRL) of solvometallurgical processes is still low. Problems that have to be solved before the application of the processes on a large scale include the higher price of organic extractants and solvents compared to water and mineral acids, the recycling of solvents and extractants at the end of the process, and finding effective but biodegradable and environmentally friendly solvents—they must have low flammability, low toxicity, and a low environmental impact.

The research on Sb recovery from nonferrous metal extraction is conducted mainly at the laboratory- and (rare) pilot-scale levels. More industrial experiments are highly needed. In addition, life cycle assessments and economic feasibility studies have to be carried out to identify the most auspicious techniques for industrial implementation.

In short, this review can serve as a basic reference for researchers interested in Sb recovery from low-grade ores and secondary resources by applying the hydrometallurgical route.