New Process for Efficient Separation and Comprehensive Recovery of Valuable Metals from Jarosite Residues

Abstract

Jarosite residue (JR), a hazardous solid waste generated in non-ferrous metallurgy, poses significant environmental challenges due to its large volume and poor storage stability. However, its high content of valuable metals (such as iron, zinc, gallium, indium, silver, …) makes its efficient recovery and comprehensive utilization highly significant. This study investigates the “oxidative roasting–reductive smelting” process for JR treatment. The reduction thermodynamics of JR-R (roasted JR) were analyzed, and the effects of smelting temperature, time, and slag basicity on the reduction and smelting process were examined. The results indicate that increasing slag basicity and temperature generally decreases slag viscosity. Thermodynamic calculations demonstrate that reductive smelting effectively enriches valuable metals (>1039 °C). The optimal conditions for reductive smelting of JR were determined to be as follows: smelting temperature of 1550 °C, smelting time of 60 min, and slag basicity of 0.9. Under these conditions, the process achieved an Fe grade of 92.87% in pig iron with a recovery rate of 90.66%, a Ga grade of 377 g/t with a recovery rate of 94.91%, and Zn and In volatilization rates of 99.91% and 83.36%, respectively. This study provides a feasible approach for the comprehensive recovery of valuable metals such as Ga, Fe, Zn, and In from JR, offering promising economic and social benefits.

1. Introduction

Jarosite residue (JR) is a solid waste produced during the iron removal process in zinc hydrometallurgy plants. It typically contains valuable metals such as Fe, Pb, Zn, and Cu, as well as trace amounts of high-value rare and precious metals like In and Ga, making it economically valuable for recovery. The residue generally consists of 25–30% Fe, 3–6% Zn, 0.5–2% Pb, and 100–300 g/t In [1,2,3,4]. According to the <<National Hazardous Waste List>> (2021 edition), JR is classified as solid waste with a waste code of 321-004-48, categorized under non-ferrous metal smelting waste. The most common method for handling jarosite residue is stockpiling [5,6,7,8]. While jarosite can remain stable under certain acidic conditions, it tends to hydrolyze or decompose when pH levels rise or under thermal conditions, leading to the leaching of heavy metal ions, which can pollute soil and water environments [9,10,11]. Therefore, the long-term storage of jarosite residue is not advisable, and timely treatment and proper utilization are essential. Otherwise, it would result in resource waste and potential environmental hazards. Direct disposal or stockpiling can lead to the leaching of heavy metal ions during natural accumulation, causing heavy metal contamination of soil and water bodies, as well as the loss of iron resources and other valuable metals [12]. Consequently, in the context of growing resource scarcity and increasingly stringent environmental regulations, developing an economically viable and environmentally friendly approach for the comprehensive treatment of large volumes of jarosite residue is a significant challenge facing the non-ferrous metallurgy industry, crucial for the sustainable development of the zinc metallurgical sector.

Currently, the disposal of jarosite residue can be categorized into two main approaches: harmless stabilization and the recovery of valuable metals. Harmless stabilization techniques for jarosite residue include high-temperature sintering, reduction roasting–magnetic separation, high-temperature hydrolysis, and solvent leaching [5,13]. However, these harmless disposal methods have not been widely adopted due to their inability to fully utilize the synergistic recovery of iron and other valuable metals, as well as their high operational costs. The existing methods for extracting valuable metals from jarosite residue mainly target specific metals, failing to achieve comprehensive utilization. Therefore, there is an urgent need to develop a method that enables the efficient recovery of valuable metal resources while being applicable to industrial-scale operations.

Significant progress has been made in the resource recovery and harmless treatment of jarosite residue, primarily reflected in the following three aspects: (1) Valuable elements such as lead, copper, cadmium, nickel, zinc, and silver can be extracted from jarosite residue using hydrometallurgical, pyrometallurgical, or combined pyro-hydrometallurgical methods, with recovery rates for certain metals exceeding 90%. (2) Jarosite residue can be used to manufacture construction materials such as cement and sintered bricks, which are then applied as cement additives in construction projects. (3) Jarosite residue can also be used in the preparation of other high-value-added materials, such as electrode materials for energy storage systems, from jarosite residue [14]. The hydrometallurgical process for metal recovery generally involves three main stages: leaching, solution concentration and purification, and metal recovery [15]. The main advantage of hydrometallurgy is its high metal recovery rate, making it an effective solution for metal extraction and suitable for processing complex and impure ores, making it a versatile option for different ore types. However, it is associated with high costs due to complex procedures and the need for chemicals, as well as challenges such as equipment corrosion, degradation, and environmental pollution [16,17]. Pyrometallurgical processes, known for their high processing efficiency and acceptable heavy metal recovery rates, have become increasingly widely adopted [18]. These methods primarily involve high-temperature roasting decomposition, where jarosite is first roasted at high temperatures to break down its crystal structure, followed by carbothermic reduction to convert it to metal. The metals are volatilized and then collected as oxides. Typically, this process is conducted at temperatures ranging from 1100 to 1300 °C, with coke or fine coal added to create a reducing atmosphere [19]. This approach is characterized by strong adaptability to raw materials, simplicity of the process, high operability, and high comprehensive metal recovery rates, making it a prevalent method for treating metallurgical waste both domestically and internationally [20,21]. However, current hydrometallurgical processes still face high reagent consumption in industrial applications [22].

In this study, jarosite residue was subjected to oxidative roasting and reduced through smelting with coke (used as a reducing agent) and flux. During the process, Zn and In in the material were reduced to metallic vapor and entered the gas phase, where they were oxidized to form oxides (ZnO and In₂O₃), ultimately concentrating in the dust. Ga, due to its strong affinity for iron, followed a similar path as iron during the reduction smelting process and accumulated in the smelted pig iron. This research provides a feasible pathway for the comprehensive recovery of dispersed metals such as In and Ga, as well as Fe and Zn resources, from jarosite residue. This approach enables the recovery of valuable metals while mitigating environmental hazards, demonstrating promising potential for both economic and social benefits.

2. Materials and Methods

2.1. Materials

The JR used in this study was sourced from a domestic smelting plant. The JR was pre-dried at 100 ± 5 °C for 10 h to remove free moisture before being subjected to chemical analysis. The chemical compositions of both the raw JR and roasted, pre-treated JR are shown in

Table 1. The chemical composition of JR/wt.%.

	Element	TFe	S	Zn	Cu	Pb	SiO2	CaO	Na2O	MgO	K2O	Ag *	Ge *	In *	Ga *	LOI
Substance
JR		28.25	12.78	3.61	0.11	0.074	0.50	0.19	1.92	0.46	0.0086	30.05	18.4	490	160	49.33
JR-R		57.15	0.37	6.07	0.22	0.016	1.15	0.11	2.30	0.98	0.20	44.05	39.2	750	220

Note: * g/t.

. (According to the corresponding national standards, chemical titration and X-ray fluorescence (XRF-PANalytical, Almelo, The Netherlands) analysis were employed for the determination of the chemical composition of JR). It can be observed that the primary valuable metals in the JR are Fe and Zn, with concentrations of 28.25% and 3.61%, respectively. The concentrations of the dispersed metals In and Ga are 490 g/t and 160 g/t, respectively, indicating significant potential for resource utilization. The contents of hazardous elements such as As and Cd are relatively low, while the loss on ignition (LOI) reaches values as high as 49.33%. According to previous studies [22,23,24,25], the decomposition of jarosite residue involves two distinct stages of significant mass loss as the temperature increases: dehydroxylation occurring between 280 and 450 °C, followed by the release of sulfur from sulfates as SO₃ gas in the range of 580 to 730 °C.

To further investigate the phase composition of the JR, X-ray diffraction (XRD) analysis was conducted on the sample. As shown in Table 1 and Figure 1, the primary phase present in the JR is sodium jarosite (NaFe₃(SO₄)₂(OH)₆). After the roasting pre-treatment, the main components of the JR are Fe₂O₃ and ZnO•Fe₂O₃.

Coke was selected as the reducing agent for the subsequent reduction smelting experiments. The coke was provided by a domestic steel plant. The results of the proximate analysis and the chemical composition of the ash are presented in

Table 2. Industrial analysis of reductant/wt.%.

Industrial Analysis	Mad	Aad	Vad	Fcad
Coke	0.59	11.98	2.44	85.58

Note: All samples are on an air-dry basis. M_ad: moisture content; A_ad: ash content; V_ad: volatile matter content; Fc_ad: fixed carbon content.

and

Table 3. Ash chemical composition of reductant/wt.%.

Chemical Composition	TFe	SiO2	CaO	MgO	Al2O3	P	S	K2O	Na2O
Coke	5.57	43.98	5.51	0.95	32.55	0.33	0.11	0.69	1.69

, respectively.

From the proximate analysis results in Table 2, it can be observed that the fixed carbon content in the coke is 85.85%, with an ash content of 11.98% and a sulfur content of 0.11%, indicating that it is high-quality coke. The main components of the coke ash are SiO₂ and Al₂O₃, as shown in Table 3.

2.2. Methods

Figure 2 illustrates the experimental flowchart for the comprehensive recovery of valuable metals from JR. The detailed experimental procedure is described below.

2.2.1. Jarosite Briquettes

The jarosite was dried to a moisture content of approximately 10%, then manually crushed to a particle size of −1 mm for later use. It was mixed uniformly in a mixer to achieve the homogenization of fractional proportions. Subsequently, a briquetting machine was used to press the material into cylindrical briquettes with a base diameter of Ø10 mm and a height of 10 mm. The briquettes were dried in a forced-air oven at 105 °C for 4 h for further use.

2.2.2. Roasting Desulfurization Test

The desulfurization roasting test of the dry briquettes was conducted in a controlled atmosphere horizontal tube furnace. Ten dried briquettes were weighed and placed into an alumina crucible. The briquettes were roasted for 20 min at 1300 °C, which allows for the rapid decomposition of sulfates [26,27] and ensures their adequate consolidation for sufficient mechanical strength; they were then rapidly removed and cooled using liquid nitrogen. The residual sulfur content in the roasted briquettes was measured, and the desulfurization rate was calculated.

The equation is as follows:

$$\mathrm{\phi }={\displaystyle \frac{m_1·S_1-m_2·S_2}{m_1·S_1}}\times 100\%$$

(R1)

where

φ = desulfurization rate of JR oxidation roasting, %;

m₁ = mass of the dry briquettes before roasting, g;

S₁ = sulfur content in the JR before roasting, %;

m₂ = mass of the briquettes after roasting, g;

S₂ = sulfur content in the briquettes after roasting, %.

2.2.3. High-Temperature Smelting Test

The experiment was conducted in an SG-6-160 resistance furnace (380 V, 12 kW, 1650 °C). The smelting container used was a graphite crucible with the following specifications: bottom diameter of Φ50 mm, top diameter of Φ65 mm, and height of 100 mm. The slag for this experiment was designed using binary basicity (R = CaO%/SiO₂%), and the amounts of roasted briquettes, coke (reducing agent), and other slag-forming agents (all of analytical grade) were weighed according to the batch calculation and mixed thoroughly. The mixed material was placed into the graphite crucible, which was then placed in the high-temperature chamber of the furnace that had reached the target temperature. The temperature was raised to the target value, and once the set temperature was reached, the heating was maintained for a specified period. After the set time, the graphite crucible was removed, covered, and buried in waste coal for cooling to room temperature. The pig iron was then removed and weighed, and the slag was collected. The contents of Fe, Ga, Zn, In, and other elements in the pig iron and slag samples were analyzed to calculate the corresponding technical parameters.

The recovery rates of Fe and Ga can be calculated as follows:

$$\mathsf{\sigma }={\displaystyle \frac{m_2·{\gamma }_2}{m_1·{\gamma }_1}}\times 100\%$$

(R2)

where

σ = Fe and Ga recovery rate, %;

m₁ = mass of the oxidized roasted briquettes used for reduction smelting, g;

γ₁ = Fe and Ga content in the oxidized roasted briquettes, wt.%;

m₂ = mass of pig iron, g;

γ₂ = Fe and Ga content in the pig iron, wt.%.

The recovery rates of Zn and In can be calculated as follows:

$$\mathsf{\mu }={\displaystyle \frac{m_1·{\alpha }_1-m_2·{\alpha }_2-m_3·{\alpha }_3}{m_1·{\alpha }_1}}\times 100\%$$

(R3)

where

μ = volatility rate of Zn and In, %;

m₁ = mass of the oxidized roasted briquettes before reduction smelting, g;

α₁ = Zn and In content in the oxidized roasted briquettes, wt.%;

m₂ = mass of pig iron, g;

α₂ = Zn and In content in the pig iron, wt.%;

m₃ = mass of slag, g;

α₃ = Zn and In content in the slag, wt.%.

3. Results

3.1. Basic Research on the Reductive Smelting Process

3.1.1. Raw Material Properties and Batch Calculation

The reduction smelting experiment was conducted using the roasted briquettes of JR. The purpose of this process is to reduce Zn and In, causing them to volatilize and enter the smoke dust while separating them from the slag, and to ensure that Ga enters the pig iron. It is assumed that the slag and iron are completely separated, meaning that Fe and Ga all enter the hot metal and that SiO₂, Al₂O₃, CaO, MgO, and Na₂O from the roasted briquettes all enter the slag phase. In this case, the contents of SiO₂, Al₂O₃, CaO, MgO, and Na₂O in the slag are as follows: SiO₂ = 18.31%, Al₂O₃ = 27.71%, CaO = 1.75%, MgO = 15.61%, and Na₂O = 36.62%, with a basicity of 0.096. The SiO₂—Al₂O₃—Na₂O phase diagram (by using FactSage software 8.0) for the fixed slag, with CaO = 1.75% and MgO = 15.61%, is shown in Figure 3. It can be seen that the melting point of this slag is approximately 1700 °C, which is too high and exceeds the requirements for conventional electric furnace smelting. Therefore, adjustments to the slag system are necessary to reduce the melting point and viscosity of the slag as much as possible to meet the requirements of conventional electric furnace smelting.

Assuming a reduction smelting test with 100 kg of roasted briquette, the calculation for the charge materials is as follows:

The required carbon amount for reducing iron oxide is calculated based on the reaction: 3C + Fe₂O₃ = 2Fe + 3CO. The carbon amount required is

$${\mathrm{M}}_1={\displaystyle \frac{100\times 0.5715\times 3\times 12}{2\times 56}}=18.36\mathrm{kg}$$

The required carbon amount for reducing zinc oxide is calculated based on the reaction: C + ZnO = Zn + CO. The required carbon amount is

$${\mathrm{M}}_2={\displaystyle \frac{100\times 6.07\%\times 12}{65}}=1.12\mathrm{kg}$$

The amount of carbon required to reduce 100 kg of roasted briquette is M₁ + M₂ = 19.48 kg. Assuming a coke utilization rate of 95%, the required amount of coke is ${\displaystyle \frac{19.48}{0.8558\times 95\%}}$ = 23.96 kg.

The masses of SiO₂, Al₂O₃, CaO, MgO and Na₂O in 100 kg of roasted briquette are as follows:

$$\begin{array}{l}{\mathrm{M}}_{\mathrm{SiO2}}=100\times 1.15\%=1.15\mathrm{kg};{\mathrm{M}}_{\mathrm{CaO}}=100\times 0.11\%=0.11\mathrm{kg}\\ {\mathrm{M}}_{\mathrm{Al2O3}}=100\times 1.74\%=1.74\mathrm{kg};{\mathrm{M}}_{\mathrm{MgO}}=100\times 0.98\%=0.98\mathrm{kg}\\ {\mathrm{M}}_{\mathrm{Na2O}}=100\times 2.30\%=2.30\mathrm{kg}\end{array}$$

The typical high furnace slag contains 15% Al₂O₃, 8.0% MgO, and a slag basicity of 1.0–1.1 [28,29,30]. Therefore, the designed slag basicity is 1.1, Al₂O₃ content is 15.0%, MgO content is 8.0%, and the slag composition is 31.55% CaO, 28.62% SiO₂, 15.0% Al₂O₃, 8.0% MgO, and 16.83% Na₂O. According to the CaO-SiO₂-Na₂O-Al₂O₃-MgO five-component slag phase diagram, the melting point of the slag composition is 1450–1500 °C (see point F2 in Figure 14). Therefore, the additional mass of SiO₂, Al₂O₃, CaO, and MgO fluxes added to 100 kg of roasted briquette (considering the SiO₂, Al₂O₃, CaO, and MgO brought in by the coke ash) is as follows: M_SiO2 = 2.76 kg, M_Al2O3 = 0.31 kg, M_CaO = 4.2 kg, M_MgO = 0.11 kg.

Apart from the melting point, viscosity is an important property of slag, which affects whether the smelting process can proceed smoothly and whether the metal can adequately settle and separate through the slag layer [31]. Therefore, the smelting process requires slag with appropriate viscosity. The liquid phase amount and viscosity of the initial slag during the smelting separation of JR briquettes at 1300–1500 °C under different basicities were calculated using the Equilib and Viscosity modules of the thermodynamic software FactSage 8.0. The Viscosity module in the software applies to fully liquid phases, whereas the initial slag at lower temperatures is not entirely liquid, necessitating correction using the Einstein–Roscoe equation, as shown below [8]:

η = η₀(1 − c)^−2.5

(R4)

where

η—viscosity of the slag, Pa∙S;

η₀—viscosity of the slag in a fully liquid phase, Pa∙S;

c—solid phase content, %.

The changes in the liquid phase amount and viscosity of the initial slag during the smelting separation of JR briquettes at 1300–1500 °C under different basicities are shown in Figure 4. As seen in Figure 4, with the increase in basicity and temperature, the slag viscosity generally shows a downward trend. Therefore, to achieve suitable fluidity, the temperature and basicity can be appropriately increased.

3.1.2. Thermodynamic Analysis of the Reduction of Roasted JR

The standard Gibbs free energy of a chemical reaction serves as the criterion for determining whether a reaction can proceed. When Δ_rG^Θ_m = 0, the reaction reaches equilibrium; when Δ_rG^Θ_m < 0, the reaction proceeds in the forward direction; and when Δ_rG^Θ_m > 0, the reaction proceeds in the reverse direction. In this study, the calculation of the standard Gibbs free energy for the relevant chemical reactions was performed using the FactSage software.

The roasted product of JR primarily contains hematite and zinc ferrite as the main phases, with the rare metals indium and gallium existing in oxide forms.

Table 4. The material melting boiling point of roasted ore.

Substance	Fe2O3	Fe3O4	Fe	ZnO	Zn	In2O3	InO	In2O	In	Ga2O3	GaO	Ga2O	Ga
Melting point/°C	1565	1594	1534	1975	420	1910~2000	n/d *	565 (sublimation)	156.63	1740	n/d	660	29.78
Boiling point/°C	3414	n/d	2750	2360	907	850	n/d	n/d	2075~2100	2800	n/d	n/d	2403

* n/d—not detected.

presents the melting and boiling points of Fe, Zn, In, and gallium oxides and their elemental forms. It can be observed that metallic iron and its oxides have higher melting and boiling points, while Zn, In, and Ga have lower melting points. Zinc is volatile, indium has a relatively low vapor pressure and is difficult to volatilize, and gallium has an extremely low vapor pressure, also making it difficult to volatilize. Through thermodynamic analysis, the possibility of reducing the oxides of iron, zinc, indium, and gallium in the roasted JR is assessed, as well as the forms of the reduction products, to reveal the reduction behavior of each metal oxide during the reduction process of the roasted JR.

(1) Reduction behavior of Fe

The reduction of iron oxides follows the principle of stepwise reduction from higher-valent oxides to lower-valent oxides. In a coal-based direct reduction system, the relationship between the standard Gibbs free energy of the iron oxide reduction reaction and temperature is shown in

Table 5. Reactions and corresponding $∆\mathrm{G}{\displaystyle \frac{\mathsf{\Theta }}{\mathrm{T}}}$ Iron oxide.

Reaction	$∆\mathbf{G}\frac{\mathsf{\Theta }}{\mathbf{T}}$ J/mol	Equation
3Fe2O3(s) + CO(g) = 2Fe3O4(s) + CO2(g)	ΔrGΘm = −52,131 − 41T	(R5)
Fe3O4(s) + 4CO(g) = 3Fe(s) + 4CO2(g)	ΔrGΘm = 3256 + 4.21T	(R6)
Fe3O4(s) + CO(g) = 3FeO(s) + CO2(g)	ΔrGΘm = 35,380 − 40.16T	(R7)
FeO(s) + CO(g) = Fe(s) + CO2(g)	ΔrGΘm = −17,883 + 21.08T	(R8)
C(s) + CO2(g) = 2CO(g)	ΔrGΘm = 166,550 − 171T	(R9)

and Figure 5.

The relationship between the equilibrium constant of the above reaction and the gaseous equilibrium components is expressed as

Δ_rG^Θ_m = −RTlnK^Θ        K^Θ = P_CO2/P_CO                P_CO = 1/(1 + K^Θ)

(R10)

Using Equations (R10), the equilibrium gaseous composition of reactions (R5) to (R9) as a function of temperature is calculated and plotted in Figure 6, alongside the equilibrium composition of the carbon gasification reaction.

As shown in Figure 6, the equilibrium curve for the reaction between Fe₂O₃ and CO is close to the x-axis, indicating that the equilibrium CO concentration, φ(CO), is approximately 0. This implies that even trace amounts of CO in the system can reduce Fe₂O₃ to Fe₃O₄. The equilibrium curve of the carbon gasification reaction intersects with the equilibrium curves of the indirect reduction reactions of Fe₃O₄ and FeO at points T_a (683 °C) and T_b (718 °C), respectively. When T < T_a, the CO concentration in the system is lower than the equilibrium CO concentration for the indirect reduction reactions of Fe₃O₄ and FeO, and Fe₃O₄ is the stable phase in the equilibrium system. Within the temperature range T_a < T < T_b, the CO concentration in the system exceeds the equilibrium CO concentration for the indirect reduction reaction of Fe₃O₄, and FeO becomes the stable phase in the equilibrium system. When T > T_b, the CO concentration in the system is higher than the equilibrium CO concentration for the indirect reduction reaction of FeO, and metallic Fe becomes the stable phase in the equilibrium system. Thus, to ensure the reduction of high-valence iron oxides to metallic iron, it is theoretically sufficient to provide an adequate amount of reductant and maintain the reaction system at a temperature above 718 °C.

(2) Reduction behavior of Zn

The melting and boiling points of zinc are 420 °C and 907 °C, respectively, while ZnO has higher melting and boiling points of 1975 °C and 2360 °C. The relationship between the standard Gibbs free energy of the zinc oxide reduction reaction and temperature is shown in

Table 6. Reactions and corresponding $∆\mathrm{G}{\displaystyle \frac{\mathsf{\Theta }}{\mathrm{T}}}$ Zinc oxide.

Reaction	$∆\mathbf{G}\frac{\mathsf{\Theta }}{\mathbf{T}}$ J/mol	Equation
ZnO·Fe2O3(s) + CO(g) = ZnO(s) + 2FeO(s) + CO2(g)	ΔrGΘm = 20,712.7 − 33.38T	(R11)
ZnO(s) + CO(g) = Zn(l) + CO2(g)	ΔrGΘm = 67,436 − 20.80T (420~907 °C)	(R12)
ZnO(s) + CO(g) = Zn(g) + CO2(g)	ΔrGΘm = 153,677 − 116T (907~1500 °C)	(R13)

and Figure 7. From the Δ_rG^Θ_m expression of reaction (R11), it is evident that the reaction can spontaneously occur at a minimum temperature of 348 °C. Within the temperature range of 420–907 °C, if ZnO is reduced to Zn, the metallic zinc will exist in a liquid state, with partial volatilization occurring as zinc vapor. When the reduction temperature exceeds 907 °C, the reduced metallic zinc will exist in a gaseous state.

The relationship between the gaseous equilibrium composition and temperature for the carbon gasification reaction and the various related reduction reactions of zinc ferrite is plotted in Figure 8.

As shown in Figure 8, zinc ferrite is difficult to reduce at temperatures below 590 °C. When the reduction temperature exceeds 590 °C, zinc ferrite is reduced to ZnO and FeO, and the reduction process becomes increasingly favorable as the temperature rises. Below 907 °C (the boiling point of zinc), ZnO cannot be reduced by CO. However, at temperatures above 907 °C, ZnO is reduced by CO to form gaseous metallic zinc, and the reduction becomes more favorable with increasing temperature.

(3) Reduction behavior of In

After high-temperature roasting of JR, indium exists in the form of In₂O₃. Given the low indium content in the roasted ore and the fact that carbon can only contact the surface of the briquette during the reduction process, In₂O₃ is mainly reduced through indirect reduction reactions. The relationship between the standard Gibbs free energy of the indium oxide reduction reaction and temperature is shown in

Table 7. Reactions and corresponding $∆\mathrm{G}{\displaystyle \frac{\mathsf{\Theta }}{\mathrm{T}}}$ Indium oxide.

Reaction	$∆\mathbf{G}\frac{\mathsf{\Theta }}{\mathbf{T}}$ J/mol	Equation
In2O3(s) + CO(g) = 2InO(s) + CO2(g)	ΔrGΘm = 1,393,990 − 339.67T	(R14)
In2O3(s) + 2CO(g) = In2O(g) + 2CO2(g)	ΔrGΘm = 275,469 − 187.89T	(R15)
In2O3(s) + 3CO(g) = 2In(l) + 3CO2(g)	ΔrGΘm = 74,941 − 57.95T	(R16)
InO(s) + CO(g) = In(l) + CO2(g)	ΔrGΘm = −657,586 + 139.37T	(R17)

.

The thermodynamic diagram plotted based on reactions (R14) to (R17) is shown in Figure 9. Analysis of Figure 9 indicates that the Gibbs free energy Δ_rG^Θ_m for reaction (R14) is greater than 0, meaning that this reaction cannot proceed in the forward direction. The minimum temperatures for the spontaneous forward reactions (R15) and (R16) are 1197 °C and 1020 °C, respectively. Therefore, during the reduction process, In₂O₃ is reduced by CO to form In₂O(g) and In(l). Studies have pointed out [32] that In₂O(g) can volatilize significantly at temperatures above 800 °C, while indium itself has a relatively low vapor pressure and is difficult to volatilize, with a vapor pressure of only 106.66 Pa at 1200 °C. However, it can be easily carried by zinc vapor into the flue gas collection system. The above analysis shows that when the reduction temperature exceeds 907 °C, ZnO is easily reduced to form zinc vapor, which facilitates the volatilization of indium.

The relationship between the partial pressures of In(g) and In₂O(g) and temperature in the In₂O₃ reduction system is expressed in Equations (R18) and (R19) [33].

lnP _In(g) = −27,723.1/T + 16.669

(R18)

lnP _In2O(g) = −56,548.0/T + 35.521

(R19)

Based on Equations (R18) and (R19), the relationship between the partial pressures of In(g) and In₂O(g) with temperature in the In-C-O system is shown in Figure 10. As depicted in Figure 10, the partial pressures of In(g) and In₂O(g) gradually increase with rising temperature. When the temperature is below 1256 °C, the partial pressure of In(g) exceeds that of In₂O(g), and the main reduction product of In₂O₃ in the system is In(g). Conversely, at temperatures above 1256 °C, the main reduction product of In₂O₃ shifts to In₂O(g).

Gallium in the roasted jarosite briquette exists in the form of oxides, and during the reduction process, carbon can only contact the surface of the briquette. The reduction of gallium oxides primarily occurs through indirect reduction by CO. The order of reducibility of gallium oxides, from easiest to most difficult, is as follows [7]: Ga₂O₃ → GaO → Ga₂O → Ga. The relationship between the standard Gibbs free energy of the indium oxide reduction reaction and temperature is shown in

Table 8. Reactions and corresponding $∆\mathrm{G}{\displaystyle \frac{\mathsf{\Theta }}{\mathrm{T}}}$ gallium oxide.

Reaction	$∆\mathbf{G}\frac{\mathsf{\Theta }}{\mathbf{T}}$ J/mol	Equation
Ga2O3(s) + CO(g) = 2GaO(s) + CO2(g)	ΔrGΘm = 1,078,770 − 373.73T	(R20)
2GaO(s) + CO(g) = Ga2O(s) + CO2(g)	ΔrGΘm = −671,111 + 174.74T	(R21)
Ga2O(s) + CO(g) = 2Ga(l) + CO2(g)	ΔrGΘm = −162,963 + 130.40T	(R22)
1/2Ga2O3(s) + CO(g) = 1/2Ga2O(s) + CO2(g)	ΔrGΘm = 203,841 − 99.41T	(R23)
1/3Ga2O3(s) + CO(g) = 2/3Ga(l) + CO2(g)	ΔrGΘm = 81,576 − 22.77T	(R24)

and Figure 11.

Based on the thermodynamic calculations for reactions (R20) to (R24), the relationship between the equilibrium gaseous composition of the carbon gasification reaction and the reduction reaction of gallium oxide with temperature is shown in Figure 12.

As seen in Figure 12, the equilibrium curve for the reaction between Ga₂O₃ and CO is close to the 100-axis, indicating that the equilibrium CO concentration, φ(CO), is approximately 100. This means that Ga₂O₃ cannot be reduced by CO to form GaO. The equilibrium curve for the carbon gasification reaction intersects with the indirect reduction reaction curves for the formation of Ga and Ga₂O from Ga₂O₃ at points T_a (1039 °C) and T_b (1110 °C), respectively. The equilibrium curve for the reduction of Ga₂O by CO lies below the equilibrium curve for the carbon gasification reaction, suggesting that Ga₂O can be reduced to Ga.

In conclusion, when T > 1039 °C, gallium oxide is reduced by CO to form liquid metallic gallium, and the reduction reaction becomes increasingly favorable as the temperature rises.

From the thermodynamic analysis of carbon reduction of Fe, Zn, In, and Ga roasted jarosite briquette, it is evident that conducting direct reduction experiments on roasted jarosite briquette can result in the reduction and volatilization of Zn and In, which become concentrated in the flue dust, while Fe and Ga are reduced and concentrated in the pig iron.

3.2. Study on the Reductive Smelting Process of JR

3.2.1. Study on the Reduction Smelting Process of JR

A single-factor experimental method was used to investigate the effects of three smelting parameters—smelting temperature, smelting time, and slag basicity—on the results of reductive smelting.

3.2.2. The Effect of Smelting Temperature on Smelting Performance

The experimental conditions were smelting time of 40 min and slag basicity R(CaO/SiO₂) = 1.1, with slag containing 15.0% Al₂O₃ and 8.0% MgO. The impact of smelting temperature on the recovery indicators of valuable metals is shown in

Table 9. Effects of smelting temperature on production indexes.

Smelting Temperature/°C	Grade of Fe/%	Recovery Rate of Fe/%	Grade of Ga/(g/t)	Recovery Rate Ga/%	Volatilization Rate of Zn/%	Volatilization Rate of In/%	Hot Metal Yield/%
1450	Incomplete separation
1500	90.28	90.49	342	89.05	99.78	82.84	57.28
1550	92.93	90.51	355	89.82	99.82	82.39	55.66

. At 1450 °C, the slag and iron did not fully separate due to insufficient heat in the system, which prevented the complete melting of the charge and resulted in high slag viscosity, hindering the separation of slag and iron. When the temperature was increased to 1500 °C and 1550 °C, slag and iron separation was achieved. At 1500 °C, the Fe and Ga contents in the hot metal were 90.28% and 342 g/t, respectively, with recovery rates of 90.49% for Fe and 89.05% for Ga. The hot metal yield was 57.28%, and the volatilization rates of Zn and In were 99.78% and 82.84%, respectively. At 1550 °C, the Fe and Ga contents in the hot metal increased to 92.93% and 355 g/t, while the recovery rates of Fe and Ga and the volatilization rates of Zn and In remained essentially unchanged, but the hot metal yield decreased to 55.66%. Considering all factors, a smelting temperature of 1550 °C was selected. Generally, for the smelting reduction process, extending the smelting time can facilitate the equilibrium of chemical reactions at the slag–iron interface, thereby improving metal recovery rates.

3.2.3. The Effect of Smelting Time on Smelting Performance

The experimental conditions were smelting temperature of 1550 °C, slag basicity R(CaO/SiO₂) = 1.1, with slag containing 15.0% Al₂O₃ and 8.0% MgO. The effect of smelting time on the recovery indicators of valuable metals is shown in

Table 10. Effects of smelting time on production indexes.

Smelting Time/min	Grade of Fe/%	Recovery Rate of Fe/%	Grade of Ga/(g/t)	Recovery Rate Ga/%	Volatilization Rate of Zn/%	Volatilization Rate of In/%	Hot Metal Yield/%
20	Incomplete separation
40	92.93	90.51	355	89.82	99.82	82.39	55.66
60	92.46	90.89	367	93.72	99.92	82.77	56.18
80	92.46	90.90	368	93.98	99.89	86.40	56.18

. It can be seen that at a smelting time of 20 min, the materials were not fully melted, and slag and iron could not be completely separated. When the smelting time was extended to 40 min, slag and iron separation were achieved, with Fe and Ga grades in the hot metal at 92.93% and 355 g/t, and Fe and Ga recovery rates at 90.51% and 89.82%, respectively. The volatilization rates of Zn and In were 99.82% and 82.39%, and the hot metal yield was 55.66%. Further extending the smelting time to 60 and 80 min increased the Ga content in hot metal to 367 g/t, with Ga recovery exceeding 93.5%, while other smelting indicators showed little change. Therefore, an optimal smelting time of 60 min was selected.

3.2.4. The Effect of Slag Basicity on Smelting Performance

Under fixed experimental conditions—smelting temperature of 1550 °C, smelting time of 60 min, and slag composition with 15.0% Al₂O₃ and 8.0% MgO—the effect of slag basicity on smelting performance was investigated. The results are shown in Figure 13.

As shown in Figure 14, within a slag basicity range of 0.9–1.3, the basicity has little impact on the smelting indicators. The iron content in the hot metal ranges from 92.20% to 92.87%, with an iron recovery rate of 90.66% to 90.99%; the Ga content ranges from 366 g/t to 377 g/t, with a Ga recovery rate of 93.72% to 95.61%; the Zn volatilization rate is 99.90%, and the In volatilization rate ranges from 82.77% to 84.87%. When the slag basicity is further increased to 1.5, the recovery indicators for valuable metals show a decline. Within an appropriate range of basicity, the slag exhibits good fluidity, facilitating the settling and separation of iron and gallium from the slag. When the slag basicity is within 0.9–1.3, the slag melting point is low (points F1, F2, and F3 in Figure 14). When the slag basicity exceeds 1.3, the melting point becomes too high (point F4 in Figure 14), reducing slag overheating and increasing viscosity, which worsens the slag–iron separation. Considering all smelting indicators, the optimal slag basicity is 0.9.

In summary, the optimal smelting conditions are a smelting temperature of 1550 °C, smelting time of 60 min, and slag basicity of 0.9, with Al₂O₃ content of 15.0% and MgO content of 8.0% in the slag. Under these conditions, the iron content in the hot metal is 92.87%, with an iron recovery rate of 90.66%; Ga content is 377 g/t, with a Ga recovery rate of 94.91%; the volatilization rates of Zn and In are 99.91% and 83.36%, respectively; and the hot metal yield is 55.79%. The chemical analysis results of the hot metal and slag are shown in

Table 11. Chemical composition of gallium pig iron/wt.%.

Fe	MFe	Ga *	In *	Ge *	Ag *	Zn	Cu	P	S	C	Si
92.87	90.99	377	203	48.01	7.85	0.0077	0.38	0.16	0.19	4.70	0.51

Note: * g/t.

and

Table 12. Chemical composition of slag/wt.%.

Fe	SiO2	Al2O3	CaO	MgO	S	Na2O	K2O	Zn	Ga *	In *	Ag *
18.18	32.52	11.62	31.70	6.85	3.78	2.96	0.048	0.0049	39.7	13.2	4.01

Note: * g/t.

.

3.3. Economic Benefit Analysis of Valuable Metal Recovery

After conducting economic and technical analysis, this process achieves a comprehensive recovery and utilization of valuable metals such as gallium, iron, zinc, and indium, with good economic benefits and promising prospects for promotion. The theoretical metal balance diagram is shown in Figure 15, from which the economic benefits of valuable metal recovery from jarosite residue can be estimated.

Ga is used in the manufacturing of the semiconductor gallium nitride, gallium arsenide, gallium phosphide, and germanium semiconductor dopants; pure gallium and low-melting alloys can serve as heat exchange media for nuclear reactions and as filling materials for high-temperature thermometers. If this process is applied to an annual production of 100 wt jarosite residue, it is estimated that 2 tons of gallium can be recovered, with an estimated economic benefit of approximately CNY 36 million.

Fe, as a key raw material in industry, is widely used in the fields of machinery, metallurgy, chemicals, and aerospace materials. If this process is applied to an annual production of 100 wt jarosite residue, it is estimated that 47 wt of iron can be recovered, with an estimated economic benefit of CNY 1.5 billion.

Zn is an important industrial metal material, primarily used in the production of hot-dip galvanized steel sheets, galvanized steel pipes, alloys, electronic components, and chemical raw materials. It has excellent corrosion resistance and electrical conductivity, making it widely used in industries such as steel, construction, electronics, and metallurgy. If this process is applied to an annual production of 100 wt jarosite residue, it is estimated that 6 wt of zinc can be recovered, with an estimated economic benefit of CNY 1.3 billion.

In, due to its strong optical permeability and conductivity, is primarily used in the production of ITO targets (used for manufacturing liquid crystal displays and flat-panel screens) and can also be applied in the electronics semiconductor field. If this process is applied to an annual production of 100 wt jarosite residue, it is estimated that 63 tons of indium can be recovered, with an estimated economic benefit of CNY 900 million.

4. Conclusions

1. Increasing the smelting temperature can increase the superheating degree of slag, reduce slag viscosity, facilitate the settlement and aggregation of metal particles, and reduce mechanical inclusion losses. Extending the smelting time helps promote the slag–iron interface reaction towards an equilibrium, thereby improving the metal recovery rate. Within an appropriate range of basicity, the good fluidity of the slag favors the separation of iron and gallium from the slag.

2. The optimal process conditions for the desulfurization roasting and reduction smelting of JR are a coke powder ratio of 23.96%, a smelting temperature of 1550 °C, a smelting time of 60 min, and a slag basicity of 0.9, where the slag contains 15.0% Al₂O₃ and 8.0% MgO. Under these conditions, the pig iron achieved an iron grade of 92.87%, with an iron recovery rate of 90.66%; the gallium grade was 377 g/t, with a gallium recovery rate of 94.91%; and the volatilization rates of Zn and In were 99.91% and 83.36%, respectively.

3. Through economic and technical analysis, this process realizes the comprehensive recovery and utilization of valuable metals such as gallium, iron, zinc, and indium, showing promising economic benefits and prospects for promotion. For an annual production of 100 wt JR, this process can recover 2 tons of Ga, 470,000 tons of Fe, 60,000 tons of Zn, and 63 tons of In, with an estimated total economic benefit of CNY 4.7 billion.