Recent Advances in Indium Recovery

Abstract

Though indium has been removed from the fifth list (2023) of critical raw materials for the European Union list of critical metals, its recovery is still of paramount importance due to its wide use in a series of high-tech industries. As its recovery is closely associated with zinc mining, the recycling of In-bearing wastes is also of interest, for both profitable and environmental reasons. With unit operations (in hydrometallurgy and pyrometallurgy or extractive metallurgy) playing a key role in the recycling of indium, the present work reviewed the most recent innovations (2024) regarding the use of these operations in the recovery from this valuable metal from different solid or liquid wastes.

1. Introduction

Though indium recently disappeared from the 2023-CRM list of the European Union [1], this metal is still considered to be of the utmost necessity for the development of a series of smart technologies for a number of reasons [2].

It is generally considered that indium has no independent deposits, though some minerals such as indite (indium–iron sulfide) contain it; thus, its production is associated to zinc mining, with China being the country which controls the vast majority of indium reserves. The majority of the global indium production comes from the by-products in the zinc refining process, where indium is enriched to varying degrees due to various smelting methods. This element is mainly concentrated in the furnace slag, flue dust, anode mud, leaching residues, and flotation tailings. To supplement or as an alternative to the above, waste indium tin oxide (ITO), solar cells, and alloys, among others, are secondary resources of indium, and its recycling is worth consideration. However, with only small amounts of the element in any given device, recovering indium from a post-consumer, i.e., LCD scrap, is considered cost prohibitive [3].

As in the case of other strategic metals, indium prices have seen relative growth over the past years (

Table 1. Variation of the indium price in the last seven years (adapted from [4]).

Year	$/kg, on 1 January
2018	380.10
2019	428.61
2020	315.11
2021	315.10
2022	477.10
2023	442.30
2024	561.60

); on 25 October 2024, the price increased to $681.34/kg [4].

This increase in price together with the current global environmental spirit, and despite the first sentence of this introduction, makes the recovery of indium for waste material a consideration for many R + D + I (research, development and innovation) groups in academia, industry, and governmental scientific institutions.

Several recent reviews considered the recycling of indium, either from LCDs [5] to CIGS thin-film photovoltaic modules [6] and waste in general [7,8]. To help achieve positive results in recycling indium, molecular recognition technology (MRT) is also recognized for its usefulness and is used to recover critical metals (including indium) from minerals and process waste [9]. Also, eco-friendly living organisms have been reported to be useful in recycling metal [10].

The present review summarizes recent investigations (2024) on the various operations involved in the recovery of indium from various sources. These sources are solid wastes or aqueous solutions (leachates), resulting from the treatment of these solid wastes, or mimic solutions resembling real ones. Among the operations involved in this recovery, leaching, adsorption, liquid–liquid extraction, etc., are considered.

2. Research Methodology

References were taken from the Scopus “www.scopus.com” and Web of Science https://webofscience.clarivate.cn/wos/woscc/basic-search databases (both accessed on 1-31 September 2024) under the following terms: indium recovery, indium and leaching, indium and liquid–liquid extraction, indium and ion exchange resins, and indium, and adsorption, for the year 2024.

3. Leaching

Various electronic devices (smartphones, monitors, and television screens) were used to recover the indium contained within them [11]. After dismantling and milling, the leaching operation was conducted at 80 °C, using 1 M sulfuric acid as a leachant solution, two hours of reaction time, and 25% In-bearing solid. As a result, nearly 90% of the indium contained in the solid source was leached, though calcium, aluminum, and iron were dissolved in the process. No data were included in the publication about the separation of the various elements from the leachate.

The next reference [12] investigated the bioleaching of indium from discarded LCD panels, both on high pulp density shredded (Sh-LCDs) and powdered (P-LCDs) materials. Two acidophilic consortiums, the mixed sulfur-iron pathways and sulfur pathways, were used in the experimentation. Indium bioleaching efficiencies through the mixed sulfur-iron pathway were 60% and 100% for Sh-LCDs and P-LCDs, respectively, with three mechanisms involved in the extraction of indium from the samples: acidolysis (reaction of aromatic acetoxy group (or other ester-leaving group) with a carboxylic acid driven by the elimination of acetic acid (or other acid)), complexolysis (reaction between the dissolved metal and the acid or chelating agent), and redoxolysis (microorganisms produce catalytic compounds which regulate the oxidation potential of the solution). In the case of the sulfur pathway, the indium recovery rates were 10% for Sh-LCD and 55% for P-LCD simple. No data about the treatment of the solution for In recovery were included in the work.

Due to their properties (adjustable solubility, high chemical stability, tunability, low cost, etc.) [13], deep eutectic solvents (DESs) are considered a type of green solvent with multiple uses. One of these uses is to act as a substitute for mineral acids in the leaching operation of various metal-bearing materials. This work [14] utilized short-chain dicarboxylic acid-choline chloride DESs to dissolve indium tin oxide, though the type of the acid has some influence on the rate of dissolution (

Table 2. Influence of dicarboxylic acids on ITO leaching.

Acid	% In Leached	% Sn Leached
Oxalic	>99	>99
Maleic	>99	>99
Succinic	59.7	70.8

Temperature: 80 °C; time: 5 h. Adapted from [14].

). The results showed that tin was leached as [Sn⁴⁺(OH⁻)₃(H₂O)]⁺ (dominant species) and [Sn⁴⁺MA²⁻(Cl⁻)₃]⁻ anion. In the case of indium, the dissolved species were [In³⁺(Cl⁻)₄]⁻ (dominant), [In³⁺MAH⁻Cl⁻)₄]²⁻, [In³⁺⁽MA²⁻)₃]³⁻,and [In³⁺MA²⁻(Cl⁻)₂]⁻. In the above formulations, MA stands for maleic acid. Tin was separated from indium by precipitation as SnO₂ (cassiterite) through dilution and hydrothermal treatment. From the resulting solution, indium was precipitated as In(OH)₃, then, the calcination of the solid at 600 °C transformed the hydroxide into In₂O₃ oxide.

The presence of lead in perovskite solar cells (PSCs) is a challenge with respect to the safe disposal of these materials [15]. In these cells, the commonly used hole transport layer is Spiro-OMeTAD, which was separated in the chlorobenzene medium. Once this layer was separated, the resulting solid was treated with aqua regia for ten minutes in order to yield a liquid phase and a solid ITO substrate, which served to fabricate new cells. Note the use of chlorobenzene: this organic solvent presented a high flammability, and though at a first instance it is considered a non-carcinogenic agent (group D) [16], other documents alleged the carcinogenic assessment of the solvent [17]. Its handling can be the cause of the development of several illnesses in humans.

The ultrasonic microwave leaching of indium from a zinc oxide dust was investigated [18]. The best conditions for the leaching operation were summarized as follows: ultrasonic power 200 W, microwave power 200 W, leaching time 30 min, sulfuric acid concentration 180 g/L, and L:S ratio 8:1. Under these conditions, the leaching rates of indium, zinc (97.69%), aluminum, (74.57%), and iron (48.62%) were reached, these results being better than those derived from the sole use of ultrasound or microwave methodologies. There were no data regarding the separation of the metals dissolved in the process.

In the treatment of zinc residues, a practical dissolution problem is often found when indium-bearing zinc ferrite (IBZF) is present in the starting material. The next reference tried to solve this problem by using a deep eutectic solvent formed by oxalic acid and choline chloride [19], with water utilized as a co-solvent and diluent. Under specific leaching conditions (temperature 90° C, liquid:solid ratio 20:1, molar ratio of oxalic acid:choline chloride 1:1, molar ratio of DES:H₂O 1:3, and reaction time 3.5 h), the leaching efficiencies of indium, zinc, and iron were 96.27%, 95.55%, and 96.33%, respectively. By the dilution of the leachate with water (leachate:water 1:4), zinc was precipitated (93.06% efficiency), and after, the ultraviolet irradiation of the supernatant produced the precipitation of iron (98.07%). Indium in the final solution was adsorbed onto activated carbon. No data were given in the manuscript about the indium-desorption step.

In the next investigation [20], a procedure was proposed for the treatment of a silver concentrate containing 3010 g/t Ag, 1.22% Pb, and 649 g/t In. The process consists of pre-oxidation followed by oxygen pressure sulfuric acid leaching. Under the optimal conditions of a sulfuric acid pre-oxidation step, concentrated sulfuric acid, temperature 60 °C, acid-to-ore mass ratio of 0.6, and 15 h, the P-O and-OH polar groups of flotation reagents on the surface of silver concentrate particles are neutralized through sulfuric acid oxidation. In the case of the oxygen pressure leaching step, the best conditions were as follows: temperature 155 °C, 160 g/L sulfuric acid, three hours of reaction time, oxygen partial pressure of 0.8 MPa, and a liquid/solid ratio of 7 mL/g. The dissolution rates were 94.8%, 98.4% and 91.1% for zinc, copper and indium, respectively, while the lead and silver grades in the leach residue increased to 2.88% and 7754 g/t, respectively. The use of sodium lignosulfonate in the oxygen pressure leaching step was mentioned, but the manuscript did not give conditions for the separation of the solubilized metals (95.1 g/L zinc, 1.23 g/L copper, 15.7 g/L total iron (0.30 g/L Fe(II)), 98.9 mg/L In in 57.8 g/L sulfuric acid medium).

The next study also considered the utilization of the oxidative pressure leaching process, but in this case a model was built on the treatment of a ZnS concentrate [21]. Using various input data derived from the literature, the model demonstrated the next efficiencies in the recovery of zinc (95.62%), sulfur (86.62%), indium (82.42%), gallium (82.55%), cadmium (85.90%), cobalt (85.57%), germanium (68.98%), and copper (62.53%), with sulfuric acid, oxygen, and lime as the main chemical inputs, while zinc electrowinning constituted over 90% of the total electricity consumption.

A thermal–biological hybrid method was utilized in the treatment of waste LCDs [22]. In the thermal treatment (20 wt% B₂O₃, 1100 °C and 70 min), metals were transferred from the glass phase to the B₂O₃ phase. After, the biometabolites of Aspergillus niger were used for the bioleaching of In, Sr, Al, and As from the thermally treated product. The best bioleaching conditions were as follows: pulp density 10 g/L, temperature 70° C, and a leaching time of two days, and the recovery rates of 82.6% Al, 70.8% As, 64.5% In, and 36.2% Sr were reached. No data were included about the separation recovery of the elements from the resulting solution.

The next reference [23] seemed to be a consequence of the results obtained from the previous reference [22]. Following a thermal treatment in very similar conditions as in [22], the resulting material was bioleached with Aspergillus niger utilizing various rate models. Kinetic evaluation indicated an overall mixed control mechanism between chemical reactivity and diffusion. Response surface methodology identified the optimal conditions as follows: temperature 70 °C and leaching time of 29 h. Under these conditions, the leaching efficiencies were as follows: 81.4% for Al, 69.1% for As, 60.0% for In, and 33.3% for Sr. As the results were very similar to those derived from [22], as in the previous reference, there was no information about the separation of the metals from the leachate.

At work, a process for recycling waste ITO targets has been investigated to recover both indium and tin [24]. Indium oxide from waste was leached using 2 M sulfuric acid and H₂O₂, at 80 °C for five hours using a L:S ratio of 4 mL:g. Under these conditions, tin remained in the same residue as SnO₂. The In-bearing solution was processed using oxalic acid, with the best conditions summarized as follows: oxalate/indium mol ratio of 1.8, 30 °C and 1 h. The precipitate (In₂(C₂O₄)₃·6H₂O) from this operation was calcined (400 °C for four hours) to yield indium oxide.

It was found [25] that in the leaching (HCl + H₂O₂ medium) of spent surface mounted device (SMD) light emitting diodes (LEDs) considerable concentrations (

Table 3. Metal concentrations in the leachate of spent LCDs.

In	Ga	Cu	Sn	Pb	Fe	Al	Si
6.4 mg/L	28.1 mg/L	29.3 g/L	3.2 g/L	1.8 g/L	0.22 g/L	0.32 g/L	0.23 g/L

Leaching solution: 3 M HCl + H₂O₂; temperature: 75 °C; time: 4 h. L:S ratio: 10. Adapted from [25].

) of accompanying metals dissolved together with indium (and gallium), thus, a replacement to the above medium is of the utmost necessity. Thus, desferrioxamine E is considered as the replacement to treat these spent SMD-LEDs. Best conditions for indium dissolution (24%) is reached at pH 7.3, 416 µM desferrioxamine E, solid:liquid ratio 5 g:L, “room temperature” (literally form the published manuscript) and one day.

This work [26] seemed to be a variation of the previous reference [25]. In this case, the source for indium was spent LCD screens, which were leached with the same desferrioxamine E chelating agent. The best conditions for optimal indium recovery (32%) were established as follows: pH: 5.5, S:L ratio 1 g:L, 25 °C, 5 mM chelating agent, and for 90 days. A comparison of these results with those of [25] showed important differences; in [25], indium recovery dropped after two days and, in [26], 90 days were needed to leach indium. Another important difference between both references is the optimal pH to leach indium; in [25], at a pH of 5.5 indium was poorly leached (<5%) and, in [26], this 5.5 value was the optimum for indium recovery.

4. Adsorption

In [25,26], the use of desferrioxamine E to recover indium from waste materials was mentioned. Since there are about 250 known siderophores, the selection of a suitable complexant for a given metal seemed somewhat unpracticable. Thus, the next reference [27] used a density functional theory-based approach to efficiently identify siderophores with increased selectivity towards target metals, i.e., indium. Thus, the binding properties of fimsbactin A towards indium are checked and compared with those of the reference siderophore desferrioxamine B (DFOB). The calculations showed that fimsbactin A presented 128% higher selectivity for indium than DFOB, whereas the experimental validation of the method indicated that fimsbactin A has up to 40% more selective indium binding than DFOB.

In the next reference [28], and in order to separate indium from iron (III) using electrospun fibers, the presence of fluorine-containing groups in the fibers were used to regulate the Lewis basicity of the N-containing fibers adsorption sites. The results demonstrated that in single metal solutions the adsorption capacity of both elements is the same, whereas in mixed In–Fe solutions, indium is loaded onto the fibers preferentially to iron. Desorption was performed in a sulfuric acid medium, the concentration of In(III) in the solution 73.9 being times higher than that of Fe(III).

The synthesis of a metal–organic framework adsorbent, MOF-808-TA (TA, tartaric acid) was reported [29]. Indium was better removed from the solution as the pH increased from 1 (1% efficiency) to 4 (near 40% efficiency). This element was removed (>80%) from solutions preferably to a series of other elements (<5%, Na, K, Ca, Mg, Co, Cu, Fe, Mo, Zn, Al, Sn) but Eu (near 75%) or Dy (near 80%). Indium can be desorbed (24 h) using 0.01 M HCl medium.

Two adsorbents, mesoporous silica (SBA-15) and ligand-modified (pyromellitic dianhydride) mesoporous silica SiO₂/PMDA, were used to investigate its performance on In removal from solutions [30]. SBA-15 material did not perform well in indium removal. Instead, the use of the modified silica allowed for the selective separation (pH 3) of indium in the presence of gallium, zinc, and tin. Desorption was investigated using 0.1 M nitric acid solutions.

The use of P507@MAC adsorbent on indium removal was investigated in the next reference [31]. This adsorbent resulted from the impregnation of P507 extractant (2-ethylhexyl 2-ethylhexyl phosphate, C₁₆H₃₅O₃P, (R(RO)POOH)) on coconut shell-derived mesoporous activated carbon. Indium was removed well from the solution at pH 1, allowing for the selective separation of this element from others (Ga, Cu, Cd, Zn). Indium desorption is affordable at moderate (3–5 M) nitric acid concentrations. The adsorbent was utilized in the separation of gallium and indium from CIGS photovoltaic modules leach solution (pH 1). After two consecutive steps, indium was completely removed from the solution, leaving gallium and other elements in it. Gallium was then removed from the solution (pH 2).

Since the recovery of indium from tail liquid zinc ore has difficulties due to the large concentrations of zinc and acid present in the solution, an adsorbent derived from the introduction of a network of zirconium glyphosine on carbon nanotubes was formed [32]. The composite (xZrGP-CNT) (x denoted the mass ratio of glyphosine to ZrCl₄) was used in the adsorption of indium from synthetic solutions. The maximum metal adsorption (around 80%) was derived from experiments carried out at pH values 3.5–4, increasing the removal percentage to around 90% when x = 2 or 3 and widening the pH range until 2. It was concluded that the mechanism of indium adsorption is due to electrostatic attraction, ion exchange reaction, and coordination interaction. Indium(III) was preferentially adsorbed (separation factors in the 1.7 value order) in the presence of Fe(II):Al(III):Zn(II):Mg(II):Ca(II) using a solution (unknown pH) with a concentration ratio of 1(In):1.5(Fe):5(Al):30(Zn):0.6(Mg):0.7(Ca). About 92% of indium was desorbed using 1 M nitric acid solution.

In the next reference [33], a nanofiber (PCA-Nanofiber) was fabricated by introducing an electronegative fluorine-containing group to modulate the electronic structure of adsorption sites used on indium recovery. In the present case, maximum indium adsorption occurred at pH 2, whereas desorption was performed using 1 M sulfuric acid.

Nanoscale metal–organic framework UiO-66-(COOH)₂ particles were embedded in sodium alginate (SA)-calcium structure to fabricate UiO-66-(COOH)₂/SA composite beads (approximate size of 3 mm) [34]. This adsorbent removes indium from solutions of pH 4, whereas 0.05 M HCl solutions were used to desorb indium from metal-loaded beads.

Though it is complex to compare the performance of the various adsorbents in terms of indium uptake, mainly due to the very different experimental conditions used in the investigations,

Table 4. Maximum indium uptakes onto different adsorbents.

Adsorbent	a[In]max, mg/g	Reference
PCA-nanofibers	135.7 (single In solution)	[28]
PCA-nanofibers	111.6 (mixed In-Fe(III) sol.)	[28]
MOF-808-TA	172.4	[29]
SBA-15	5.2	[30]
SiO2-PMDA	38.0	[30]
2ZrGP-CNT	193.7	[32]
PCA-nanofibers	137.9	[33]
AM-nanofiber	43.1	[33]

^a Values from the Langmuir isotherm model.

summarizes the different maximum indium loadings on selected adsorbents. With greater discrepancies between the results presented in [30] and the other references, it is worth noting here the close similarity in the adsorbents used in [28] (published online 21 December 2023) and [33] (published online 26 November 2023) and also in the maximum indium uptakes.

5. Liquid–Liquid Extraction

Crown ethers (CEs) with varied cavity sizes, benzo-18-crown-6 (B₁₈C₆), benzo-15-crown-5 (B₁₅C₅), and benzo-12-crown-4 (B₁₂C₄), were synthesized [35] and used in the extraction of indium from a KCl medium [36]. With InCl₄⁻ being the predominant metal-species in the aqueous phase, the extraction order followed the sequence B₁₈C₆ > B₁₅C₅ > B₁₂C₄. The metal extraction was attributed to the coordination of the crown ether with K⁺ to form [CE-K]⁺ at the two-phase interface, which further associates with InCl₄⁻ to create the complex of CE-KInCl₄. Moreover, B₁₈C₆ showed excellent selectivity toward In³⁺ over competing ions such as Fe³⁺, Al³⁺, Zn²⁺, Sn²⁺ and Ca²⁺ in a mixed system. Indium was recovered (98.1% efficiency) from the loaded organic phase by using 1 M HCl as a strippant. Dichloromethane was used to dissolve the extractants and this misuse or poor choice of the diluent clearly precluded the practical use of these systems due to the harmful characteristics of this chemical [16,17].

Due to their properties, ionic liquids in general and particularly phosphonium-based ionic liquids are being considered as extractants for many metals. Cyphos IL104 (trihexyl(tetradecyl)phosphonium bis(2,4,4-trimethylpentyl)phosphinate dissolved in toluene was used to extract indium from a feed solution derived from the leaching of discarded LCD screens [37,38]. From the In-Sn-loaded organic solution, indium was selectively stripped with 4 M nitric acid. In a leaching-extraction procedure, the ionic liquid [P₄₄₄₁₀][Br₃] (tributyldecylphosphonium bromide) was used to recover indium (and gallium) from GaAs and InAs semiconductors [38,39]. After this first step, arsenic was first recovered by stripping with 4 M NaBr, then gallium was stripped using water, and finally indium was recovered from the organic phase using 0.25–1.3 M NaOH solutions as strippants.

Organic phases of D2EHPA (di(2-ethylhexyl)phosphoric acid) and 2-ethylhexanol in kerosene or petroleum paraffins were used to extract indium [40]. The optimum organic phase composition to extract this metal contained 15.8% extractant and 4.7% alcohol. In this extraction stage, indium recovery was in the 97.5–99.5 range. Mixtures of sulfuric acid and sodium chloride were used as strippants with efficiencies of around 99%. Though in the Abstract it is mentioned that the introduction of the alcohol into the organic phase helps to improve the indium “re-extraction” compared to D2EHPA alone, it is more than possible that the use of the alcohol prevented the formation of a third phase (or second organic phase) after the disengagement phase either in the extraction and/or stripping stages. The main problem with using D2EHPA to extract indium is the co-extraction of iron(III) [41]

The extraction of different technologically critical elements from leach solutions of multiple sources using ionic liquids was reviewed [42]. Generally, these extractants have low selectivity toward Ga, Ge, and In compared to Fe(III), Al(III), and Zn(II), which are typically present at high concentrations in the multielemental solutions.

Spent liquid crystal display panels were used as a source for indium recovery. The treatment of these panels [43] led to a solution which was processed using an ethanol-mediated dodecylamine (DDA) phase to which n-hexane was added prior to the separation phase [44]. Indium extraction (about 90%) occurred at pH values in the 7.5–9 range; however, tin, aluminum, and iron were extracted within the same efficiency, as well as copper (80%). In(III) loaded in the organic phase was then converted into InOOH nanoparticles with “abundant” oxygen vacancies through a low-temperature hydrothermal process. The nanoparticles (polyhedral morphology with average size 15.4 nm) were used to electrocatalyze CO₂ reduction.

6. Miscellaneous Operations

Besides the most used unit operations in the recovery of indium and reviewed above. Scientists try to resolve the problem using different methodologies. The results are reviewed below.

Ultrasound-enhanced calcium carbonate precipitation was used to increase indium recovery on a zinc oxide dust leachate (68 g/L Zn(II), 0.56 g/L In(III), 1.5 g/L Fe(II), 0.4 g/L Al(III)) [45]. The results of this investigation showed that the best conditions were as follows: a precipitation endpoint pH of 4.0 and an ultrasound power of 200 W. The precipitation yield of indium under these conditions was 99.79%. Compared with conventional precipitation, the utilization of ultrasounds reduced both the time to reach equilibrium, and the amount of calcium carbonate used in the process. It should be noted that together with indium, aluminum was precipitated within the same rate order (96.75%), whereas the rate of zinc precipitation reached 29.36%. Thus, the precipitated In(OH)₃ may be contaminated to some extent by these two elements.

Membranes, and particularly supported liquid membranes, are a widely proposed methodology in the removal of metals from liquid solutions or effluents [46]. In this methodology, an extractant or an organic phase (extractant and diluent as used in liquid–liquid extraction) are immobilized in the micropores of a hydrophobic solid support, forming the supported liquid membrane. Indium was recovered from an aqueous solution using one of these liquid membranes, acting Cyphos IL104 diluted in kerosene as the organic phase for the transport process [47]. The following five process parameters were evaluated: indium concentration (10–100 mg/L), extractant concentration (0.05–0.2 M), feed phase acidity (0.01–3 M nitric acid), chloride ion concentration (0.5–4 M), and the stripping agent concentration (0.1–5 M nitric acid). Experimental results were compared with the models resulting in the utilization of the response surface methodology (RSM with r² = 0.9589) and artificial neural network (ANN with r² = 0.9860)). The best-optimized data were as follows: 73.92 mg/L (In), 0.157 M (Cyphos IL104), 1.386 M (nitric acid in the feed phase), 2.99 M (chloride ion), and 3.06 M (nitric acid in the stripping phase).

Copper indium gallium selenide solar cells were used as the source phase for indium recovery using microwave pyrolysis, thermal oxidation, and thermal chlorination technologies [48]. Under microwave pyrolysis conditions (332–430 °C and 200–300 W), the cells lost about 90% of their weight. Selenium was then recovered using thermal oxidation at 800 °C for 30 min; this time was increased to 40 min if the temperature decreased to 600 °C. Further, a first thermal chlorination step (two hours at 260 °C) allowed the recovery of gallium (56% efficiency with 100% pure gallium), and a second step (two hours at 340 °C) allowed the recovery of gallium (40% efficiency and 20% purity) and indium (46% efficiency and 55% purity). Gallium can be completely recovered in the first stage if longer reaction times were used. If this option was chosen, the purity of indium would rate near 68%, this being an unsatisfactory result. To increase this rate, the usage of one or more additional technologies is necessary to achieve the complete separation of indium and other metals.

Gallium-based liquid metal (G-LM) waste is produced during the preparation of liquid metals and related products, this being a waste source of gallium and indium. In the next investigation [49], a two-stage vacuum distillation method was used to recover Ga–In alloy from these wastes. Firstly, low-temperature vacuum distillation was used to recover 98.23% of Zn under optimal conditions (1–5 Pa, 500 °C, one hour). Secondly, the Ga–In alloy was recovered through high-temperature vacuum distillation. Under the optimal conditions (1–5 Pa; 1250 °C, two hours), the recovery rates of gallium and indium were 85.14 and 95.23%, and the content of the Ga–In alloy in the product was 99.46%.

The next study [50] presents evaporation parameters and the characteristics of indium in the process of vacuum distillation to direct the refining of this metal using vacuum distillation. The investigation was carried out on a 92In–8Sn alloy. The following was concluded: (i) the measured volatilization rate of indium is temperature-dependent in the 900–1100 °C range (5 Pa and 25 mm crucible depth); (ii) the volatilization rate of indium in the alloy was found to be proportional to the crucible depth (25–65 mm at 1050 °C and 5 Pa); (iii) the determined mass transfer coefficient for indium showed positive and negative correlations with temperature and crucible depth, respectively; and (iv) a kinetic model for indium volatilization was derived and showed that the volatilization of indium in the alloy is mainly constrained by the transfer of mass in the vapor phase.

As is widely known, end-of-life LEDs are a source for the recovery of indium or rare earth elements. The next investigation [51] used these LEDs to recover indium by a prior thermochemical step with hydrochloric acid-containing vapor (650 °C) from polyvinyl chloride as a chlorine source. The results from this operation showed that indium is mainly concentrated (42–70 ppm) in liquid products, especially in the water phase. Indium can be recovered from the aqueous phase by electrolytic extraction after its concentration has been increased by reduction or by the removal of water [52]. Further, the LED residue is leached with sulfuric acid (32%) and hydrogen peroxide in a L:S ratio of 10. From the leachate, the heavy rare earth element lutetium can be extracted tributyl phosphate/1-octanol.

The infamous copper indium gallium selenide solar cells were used for the recovery of silver and indium tin oxide particles [53]. The process to achieve this used two-step ultrasonic leaching with 0.1 M nitric acid solutions. In the first step, indium tin oxide particles were liberated through the selective dissolution of the zinc-rich layer underneath, using low ultrasound power for 3 min. In the second step, the same conditions as in the first step were applied, but now a high-power ultrasound was used for 15 min to recover the silver grid lines. Both the indium tin oxide (70.5 wt% purity) and the silver grid (95.0 wt% purity) particles were recovered by filtration, and there was no loss of silver in the leachates, though some indium was dissolved, together with zinc, in the first step. Copper indium gallium selenide material remained in the substrate of the second step and some losses of this material were produced along the various steps of the process.

Scraps from indium phosphide (InP) semiconductors are another source for indium. However, the recycling of these INP scraps (78.2% In, 18.3% P) presented some challenges mainly due to phosphorus flammability and the release of toxic and irritating smokes during combustion. To resolve these challenges, the next investigation proposed a “controlled-pressure pyrolysis–spray condensation” process [54]. Firstly, InP is decomposed (1050 °C, micronegative pressure for five hours) into phosphorus and indium; phosphorus then evaporates into the vapor phase and is sprayed with warm water (55 °C) for condensation in this same medium. Industrial α-yellow phosphorus with a purity greater than 98 wt% (recovery rates exceeding 90%) was obtained and stored in water. From the first step, an indium ingot (99.59 wt% purity) was recovered (98.2% efficiency rate).

7. Conclusions

Though nowadays the EU does not consider indium as a critical metal, the recovery of this metal from different wastes is still conducted. Apparently, there is not a consolidated market for recycling some (if not all) of these wastes with a specific emphasis on indium recovery.

In the leaching step, future research should focus on enhancing the amplification technology of ultrasonic microwave leaching equipment. Also, the utilization of siderophores in eco-friendly technologies should be taken into consideration for the recovery of indium (and many different critical metals) from various industrial waters and leachates or bioremediation approaches. The potential uses of ionic liquids and deep eutectic solvents to substitute more aggressive acidic compounds in the leaching operation cannot be forgotten. As substitutes to these mineral acids, the use of organic acids may be considered.

Adsorption processing is a technology to be considered, even though a series of adsorbents tend to lose their initial adsorption properties under continuous use; thus, one of the future hot spots in this field should be the obtention of more stable adsorbents under more extreme working conditions, i.e., acidities of the feed or desorption solutions.

In the case of liquid–liquid extraction, the rational design of ionic liquids and task-specific ionic liquids, their recycling, process scaling-up, and detailed economic viability studies need to be adequately investigated to accelerate the potential commercialization of these chemicals in the processing of indium.

Of course, the development of alternative technologies to the established ones is always of interest.

In any case, the real value of all the above technologies must be proven in the recovery of indium from real waste solid materials, or in the treatment of real leach solutions, and under continuous use (pilot plant); thus, efficiency in the recovery and selectivity of indium must be areas to consider.