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Article

Thermodynamic Study Proposal of Processing By-Product Containing Au, Ag, Cu and Fe Sulfides from Antimony Ore Treatment

by
Dušan Oráč
1,
Martina Laubertová
2,*,
František Molnár
3,
Jakub Klimko
1,
Vladimír Marcinov
1 and
Jana Pirošková
1
1
Institute of Recycling and Environmental Technologies, Faculty of Materials, Metallurgy and Recycling, Technical University of Košice, Letná 1/9, 042 00 Košice, Slovakia
2
Institute of Earth Resources, Faculty of Mining, Ecology, Process Control and Geotechnologies, Technical University of Košice, Letná 1/9, 042 00 Košice, Slovakia
3
Slovak Iron Route, Letná 9, 042 00 Košice, Slovakia
*
Author to whom correspondence should be addressed.
Processes 2025, 13(3), 842; https://doi.org/10.3390/pr13030842
Submission received: 31 January 2025 / Revised: 8 March 2025 / Accepted: 10 March 2025 / Published: 13 March 2025
(This article belongs to the Special Issue Non-ferrous Metal Metallurgy and Its Cleaner Production)
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Abstract

:
A possible thermodynamic study of processing Cu (Ag, Au) and Fe sulfide concentrate as a by-product after the processing of tetrahedrite concentrate, applying pyrometallurgical and hydrometallurgical methods, was studied. The sample of sulfide concentrate, 34.7 wt. % Cu, 21.4% Fe, 12 g/t Au, and 7.317 g/t Ag was contained. Analytical technique AAS was used to analyze the sample before conducting a thermodynamic study of the leaching of sulfide concentrate by applying Pourbaix Eh–pH diagrams. The outcome of this thermodynamic research will provide essential data to support recent hydrometallurgical technologies. If its correctness can be verified experimentally, this result will be promoted to developing a new alternative copper-production technology. The minor components Sb, As, Hg, and Bi are also present in the concentrate in the form of sulfides Sb2S3, As2S3, Bi2S3, and HgS. This theoretical proposed hydrometallurgical technology shows that it is possible to obtain Fe in the form of Fe(OH)3, and after its thermal decomposition, it can be prepared as Fe2O3 as a marketable product. In any case, the most economically advantageous would be complete hydrometallurgical processing, i.e., also Cu(Ag,Au)Fe sulfide concentrate, with the possibility of valorizing Cu, Ag, and Au in metallic form.

1. Introduction

The European Union (EU) aims to reduce the dependence on imports of raw materials, which are considered crucial for their industries. Europe is very rich in aggregates and industrial minerals, as well as some base metals such as copper and zinc [1,2]. Antimony is one of the critical raw materials the EU mainly uses in lead-acid batteries and plastic catalysts. Globally, the primary production of antimony is now isolated to a few countries and is dominated by China. China, India, Vietnam, and Tajikistan are commercial intermediaries of global antimony resources. Globally, antimony reserves are limited and unevenly distributed. China, Russia, and Bolivia account for about 80% of the global antimony reserves. Antimony reserves are located in six European countries, i.e., France, Germany, Sweden, Finland, Slovakia, and Greece. The potential future mining of antimony resources includes either simple stibnite or precious metal deposits associated with copper, lead, and/or zinc and gold [3,4,5,6].
Primary antimony is produced from stibnite ore into product antimony oxide (Sb2O3) and antimony metal (Sb) by pyrometallurgical and hydrometallurgical methods [7,8]. At this moment, mainly a pyrometallurgical process is used, where antimony metal, SO2 gas, and a slag are produced. Hydrometallurgical treatment of stibnite ore is based on two steps: alkaline sulfide leaching and acidic chloride leaching, followed by electrodeposition of antimony metal or hydrolysis with NaOH or NH4OH to produce Sb2O3. Stibnite Sb2S3 ore can be treated with a variety of leaching agents: a mixture of hydrochloric acid and tartaric acid, a mixture of nitric acid and tartaric acid, and hot concentrated sulfuric acid. Sb2S3 can also be dissolved by an alkaline solution consisting of sodium sulfide (Na2S) and sodium hydroxide (NaOH). The development of alternative methods may prove to be one of the most helpful factors for the metallurgical industry [9,10,11,12,13]. As an alternative to roasting and pressure leaching, the use of alkaline pretreatment is recommended to decompose the sulfides and release the gold. Slovakia has non-balanced reserves of antimony minerals in Dúbrava, Pezinok, Špania Dolina, Rožňava, and Rudňany deposits, with grade 0.4–2.2% of Sb. Mineralization is represented by quartz veins and stockworks with berthierite (FeSb2S4), stibnite (Sb2S3), valentinite (Sb2O3), kermesite (Sb2S2O), tetrahedrite ((Cu, Fe, Zn, Ag)12Sb4S13), and schafarzikite (FeSb2O4) [14,15,16,17].
In Slovakia, antimony was produced by the pyrometallurgical method by processing domestic (Sb2O3) concentrates with a high content of precious metals (Au, Ag). Tetrahedrite is one of the most common sulfide minerals and represents the most important source of copper and antimony, silver, and mercury. The tetrahedrite concentrates from the Rožňava deposit were produced by nonselective sulfide flotation and have a high content of copper (27%), antimony (16%), and silver (4000 g/t) [18,19,20].
The mineral is most often used as a source of copper, but the content of antimony, arsenic, and mercury, which is usually quite high, makes the tetrahedrite unsuitable for pyrometallurgical processing for both ecological and technological reasons. Hydrometallurgical methods have better applicability on pure ores, selective metal recovery, and better control of the process [21]. Tetrahedrite mineral is very refractory to the most common leaching processes, and to cause significant dissolution, it is necessary to use high temperatures and pressures [22,23].
In acid oxidative leaching in a medium of Fe3+ ions, copper and iron pass into the solution and, depending on the leaching conditions, also antimony. Alkaline leaching in sodium sulfide medium selectively dissolves the antimony, while copper and iron remain in the solid residue [24,25]. Leaching is a key step in the whole hydrometallurgical treatment scheme. This can be influenced by the selection of a suitable leaching agent (ferric chloride [26], thiourea [27], sulfuric acid [28], sodium sulfide [29]), suitable ore treatment (mechanically and mechano-chemical activation [30]), and the use of an intensification method (ozone [31], ultrasound [32], microwave radiation [33], and using deep eutectic solvents [34]). During alkaline tetrahedrite dissolution, copper is precipitated in the form of insoluble covellite and chalcocite minerals and surface chemical reaction as the rate controlling step with Ea = 61.94 kJ [35].
Intensive research into the influence of mechanical pretreatment on the leaching process of tetrahedrite concentrate from Roznava has been observed in recent years [36]. The principle of mechanochemical leaching is used to extract antimony from the concentrate into the liquid extract. The increased reactivity of mechanically activated sulfides causes a reduction in the activation energy and an increase in the leaching rate of the individual useful metals from minerals [37]. Product development with mineral processing is important, such as drilling, milling, calcination, etc. [38]. Using a continuous mill in conjunction with a chemical reactor, it was possible to extract Sb and As utilizing further alkaline leaching, as well as the Hg into the liquid extract [39].
Antimony is selectively dissolved during alkaline leaching in a sodium sulfide medium. Copper and iron remain in the solid residue after leaching. Leaching of tetrahedrite with sodium sulfide and its mechanism can be described according to Equations (1)–(3) [39]:
C u 2 S . S b 2 S 3 ( s ) + N a 2 S ( l ) = C u 2 S ( s ) + 2 N a S b S 2 ( l ) ,
N a S b S 2 ( l ) + N a 2 S ( l ) = N a 3 S b S 3 ( l ) ,
( x − 1 ) N a 3 S b S 3 l + N a 2 S x l = x − 1 N a 3 S b S 4 ( l ) + N a 2 S ( l ) ,
where antimony passes into a trivalent to pentavalent soluble complex form and copper together with silver remains in a solid form [39]. The solid residue after leaching contains sulfides of Cu, Au, and Ag, also contains Fe, and has minor impurities of Sb, Hg, As, and Bi that have not completely dissolved. This sulfide concentrate Cu (Ag, Au) as a by-product of mechano-chemical alkaline leaching of tetrahedrite concentrate can already be processed by pyrometallurgical pretreatment or hydrometallurgical treatment. Higher economic benefits can be expected in the case of hydrometallurgical recovery of these metals, as they are less economically demanding and more environmentally acceptable. The pyrometallurgical method is more energy-intensive [40].
The aim of this study was to analyze the theoretical possibilities of processing sulfide concentrate containing Fe, Cu, Ag, Au and accompanying impurities Bi and As from the treatment of antimony ores by pyrometallurgical and hydrometallurgical methods using a thermodynamic study according to the authors’ previous works.

2. Materials and Methods

The chemical analysis of the sulfide concentrate sample was carried out by the AAS method using Varian AA240+ or Thermo Scientific (iCE 3000 series, London, UK). The results from the AAS (atomic absorption spectroscopy) analysis are acceptable with a relative standard deviation value below 6% for three replicate measurements. The chemical composition of Cu(Ag,Au),Fe sulfide concentrate (solid residue after mechanochemical leaching) is shown in Table 1.
Mineralogical study of sulfide concentrate showed that elements Cu and Fe are predominantly presented in the sample concentrate in sulfide phases such as Covellite CuS, Chalcocite Cu2S, Chalcopyrite CuFeS2, Troilite FeS, and Pyrite FeS2. The minor components Sb, As, Hg, and Bi are also presented in the concentrate in the form of sulfides Stibnite Sb2S3, Orpiment As2S3, Bismuthinite Bi2S3, and Cinnabarite HgS [40].
Based on the change in the standard Gibbs energy (symbol ΔG°, measured in kJ), the predicted chemical reactions of chemical treatment of Cu(Ag,Au),Fe sulfide concentrate in NaOH solution, cementation, and thermal decomposition (Table 2, Table 3, Table 4 and Table 5) of these compounds (oxides, sulfides, and sulfates) at two considered temperatures (20 °C and 80 °C) were calculated using HSC Chemistry v 6.1 software [41].

3. Results and Discussion

Figure 1 shows the technology of complex processing of tetrahedrite concentrate, where sodium hexahydroxy antimonate Na[Sb(OH)6] is produced in the first step after alkaline leaching, which is saleable and has a higher final value than pure Sb. Subsequently, the leachate is processed after cementation by diaphragm electrolysis, and a sulfide concentrate containing gold and silver is obtained.
In the processing of Cu(Ag,Au),Fe sulfide concentrate by pyrometallurgical (Figure 2) as well as hydrometallurgical technology, the first technological step is the conversion of metal sulfides to oxides. The treatment of Cu(Ag,Au),Fe sulfide concentrate by pyrometallurgical and hydrometallurgical methods is discussed in this part. A thermodynamical study of the chemical treatment of Cu(Ag,Au),Fe sulfide concentrate by oxidative leaching in NaOH solution and also reductive roasting of the concentrate obtained after conversion was realized.

3.1. Pretreatment of Cu(Ag,Au),Fe Sulfide Concentrate

3.1.1. Chemical Treatment of Cu(Ag,Au),Fe Sulfide Concentrate in NaOH Solution

The aim of the chemical treatment of the sulfide concentrate is the conversion of sulfides to oxides. Elements Cu and Fe are present in the concentrate in sulfide phases, namely CuS, Cu2S, CuFeS2, FeS, and FeS2. The minor components Sb, As, Hg, and Bi are also present in the concentrate in the form of sulfides Sb2S3, As2S3, Bi2S3, and HgS. The thermodynamic efficiency of all the major as well as minor sulfides in a strongly oxidizing environment is practically 100%, and sulfur is also converted to Na2SO4 with 100% thermodynamic efficiency. This is due, under unchanged conditions, to the excess of the starting components 2.2 mol nNaOH, 2.2 mol nO2(g), and a temperature of 60 °C. The conversion of the individual sulfides to oxides follows chemical reactions (4) to (21). From the thermodynamic study, it is evident that all the chemical reactions are temporary and are thus shifted to the product formation side as oxides of CuO, Cu2O, Fe2O3, Fe3O4, As2O3, As2O5, Sb2O3, Sb2O5, Bi2O3, and HgO. Gibbs energy (G), its change in terms of a change in entropy (ΔS), and enthalpy (ΔH), is written as ΔG = ΔH-TΔS. All the reactions listed have a probability of proceeding at a temperature of 20 °C and 80 °C, the results of the thermodynamic calculations of the individual reactions are given in Table 2.
Eh-pH diagrams (shown in Figure 3A–C) were constructed for the sulfide pretreatment method, which confirm that Cu, Fe, and Bi in the water stability region of an alkaline environment and oxidizing conditions form the phases of Cu(OH)2, Fe2O3, and Bi2O3. These phases are characterized by low solubility and do not leach directly into the alkaline solution, but the unwanted sulfur is separated from the input raw material phases in this step. In the case of arsenic, according to the diagrams (Figure 3D), leaching of the arsenic present may occur to form arsenates A s O 4 3 − . Produced arsenates react with other metal cations present, such as Fe, which also form insoluble compounds. These insoluble compounds will proceed further in the process together with other solid phases into the roasting process described in Section 3.1.2.
From the economic and ecological points of view, it is necessary to develop a technological conversion process that allows NaOH to be recycled back into the process according to reaction (21). The calculated value of the standard Gibbs energy change indicates that the return of NaOH to the technological process is thermodynamically feasible (Figure 2 and Table 2). In addition to the NaOH solution that would be returned to the conversion process (Figure 2), BaSO4(s) in the solid phase as a marketable final product was obtained. At this stage, sulfur was not transferred in the gaseous form to the atmosphere, nor in the ionic form to water. Sulfur was not considered a pollutant in this process, which is very significant in the current increased legislative environmental requirements. Solid-phase sulfur could be considered a marketable product.
The first results of experimental verification of the conversion showed that the conversion efficiency of sulfur to sulfate of 83 to 85% was achieved [39]. After conversion, an average of 3.8% residual sulfur remained in the final product after roasting. Based on phase analysis, the conversion would produce CuO and Fe2O3. Given that the powder after conversion was strongly magnetic, it is likely to be Fe2O3-γ maghemite. The relatively high residual sulfur content of 3.8% in the powder after leaching and the relatively low conversion efficiency of S to Na2SO4 were due to the absence of mechanical activation. Mechanical activation was carried out by intensive grinding just before chemical conversion. After applying mechanical activation of the initial concentrate (by grinding) just before conversion, the residual sulfur content of the leachate (without washing) was 0.88%, and subsequently (after a single wash), the sulfur content decreased to 0.075%; thus, the sulfur conversion efficiency from 85% to 99.6% was achieved [39].

3.1.2. Treatment of the Concentrate Obtained After Conversion

First of all, it should be mentioned that the pyrometallurgical treatment of Cu(Ag,Au),Fe sulfide concentrate was also carried out on a laboratory scale. First, the conversion of Cu(Ag,Au),Fe sulfide concentrate to oxides by roasting was carried out. The purpose of solid-phase reduction roasting was to reduce only copper from the mixture of metal oxides together with the noble metals gold and silver and then convert them into a single metal phase. The optimum treatment temperature was 800 °C.
At this temperature, the reduction of CuO to Cu was carried out with optimal efficiency. From the results of the X-ray phase analysis, it was found that in addition to the Cu metal phase, the Fe metal phase, FeO, and Fe3O4 oxides were also present in the roaster in minor amounts. The minor oxides Bi2O3, Sb2O3, Sb2O5, As2O3, As2O5, HgO, and partially Fe2O3 were also reduced to elemental metal at this temperature. Separation of metallic copper can be accomplished by flotation in a Na2S environment with very good metallic copper recoveries of 90 to 92%. During the reduction roasting at 800 °C, it must be considered that sintering of the mixture of the final product of the reduction roasting process occurs, which is reflected in the coarsening of the particles.
Sintering of the reduced copper with iron may also occur, and this may cause problems in the separation of the product from the reduction roasting process. The setting conditions in the reduction roasting process of the above flotation concentrate must be carried out in an inert atmosphere to avoid oxidation of the metallic copper of the flotation grain size. During reduction roasting, the minor components, namely Sb, As, Bi, Hg, and partly Fe, are also reduced. At a roasting temperature of 800 °C, the mercury will volatilize as vapor from the system with the outgoing gases. The reduced components Sb, As, Bi, and, to some extent, Fe will dissolve and pollute the copper during smelting. Given the facts mentioned previously, copper smelting should be combined with oxidative refining of copper in a medium-frequency induction furnace under an inert atmosphere of nitrogen. For oxidative refining, copper (Cu2O) forging can be used as a source of oxygen for the transfer of impurities in the form of oxides to the slag. Oxidative refining of copper impurities using copper forgings can be described according to the following chemical reactions (22) to (25) (Table 3).
The advantage of using an induction medium-frequency furnace is that the electrodynamic forces mix the melt and thus result in a homogeneous distribution of the metals throughout the melt volume. On the other hand, the disadvantage is that the copper melt after oxidative refining would have to be poured into a non-induction furnace, where the oxides of impurities would float to the surface of the melt to form slag. Only then would it be possible to cast the copper into anodes. The production of refined copper would be carried out as usual by electrolytic refining. The noble metals Ag, Au, and Pt, which accompany copper throughout its production technology, would be introduced into the anode sludge as they always are.
The advantage of this method of processing the Cu(Ag,Au),Fe concentrate is the possibility of obtaining refined copper, and at the same time, noble metals in the anode sludge. The disadvantage of this processing method is that sintering of the reduced Cu with Fe occurs, which can cause major problems in the sintering of reduction roasting products [42].
The fine-grained particles of the Fe and Cu components are relatively homogeneously “intergrown” and associated with each other, which is confirmed by the high yield of the non-magnetic Cu component (about 80%) in the magnetic product. Even in laboratory flotation, the Cu component was floated into the foam product with only about 14% yield, with the bulk remaining in the waste after flotation. Considering the approximately equal values of the Fe and Cu component recoveries to the froth products in the flotation case, the assumption of mutual coupling of these two components, as mentioned in the reduction roasting products, is confirmed.

3.2. Treatment of Cu(Ag,Au),Fe Ssulfide Concentrate by Hydrometallurgical Technology

Leaching of Metal Oxides After Sulfide Conversion

In the processing of Cu(Ag,Au)Fe sulfide concentrate by hydrometallurgical technology, the first technological step was also, as in the case of pyrometallurgical technology, the conversion of metal sulfides to oxides, as is shown in Figure 4.
The conversion of copper oxides into solution by acid leaching using sulfuric acid solution (H2SO4) can be written according to chemical reactions (26) and (27) (Table 4).
The conversion of CuO and Cu2O oxides into solution using H2SO4 is a thermodynamically feasible solution at 20 °C. The conversion of Cu2O into solution (reaction 27) theoretically produces 50% of CuSO4, and the remaining part of copper is reduced to metallic Cu powder. Equations (30) and (31) show the leaching of Fe2O3 and Fe3O4 phases. The standard Gibbs energy change calculations confirms that these reactions are also thermodynamically possible (Table 4).
The Eh-pH diagrams shown in Figure 3A and Figure 3B show the leaching range of copper and iron, respectively. In the case of copper, it is calculated that in a thermodynamic system, Cu2+ ions are formed at a Cu concentration of 1 mol/dm3 in the range below pH 2.5. In the case of iron, much depends on the oxidation conditions of the system because Fe2+ ions are leached in a wider pH range than Fe3+ ions, which are leached at a concentration of 1 mol/dm3 only below pH 1.4. In practice, the leaching ranges shift, and the solubility increases with decreasing metal concentration in the solution.
After leaching, the next technological step is filtration. Filtration separates the residue after leaching that did not pass into the solution, namely SiO2, or other complex SiO2 compounds and noble metals Ag and Au. CuSO4 and FeSO4 are essentially passed into solution.

3.3. Precious Metals Issues in Concentrate Processing

After the conversion of metal sulfides to oxides by the action of the components of sodium hydroxide (NaOH) and oxygen (O2) solution, the conditions for the formation of Cu and Fe oxides are created with practically 100% efficiency. At the same time, conditions are also created for the formation of silver oxide (Ag2O); according to Equation (32), gold and platinum do not form such stable oxides. The above reaction is also thermodynamically viable (Table 5). Leaching of CuO, and Fe2O3 oxides in H2SO4 solution would also leach Ag2O to form silver sulfate (Ag2SO4) (reaction 37). In a subsequent step in the process of cementing copper with Fe powder, such a process would also result in the cementation of silver; this is not preferred. To obtain metallic silver in the next process step by the smelting process, it is necessary to convert Ag2O to the elemental form by thermal decomposition according to Equation (34). From a thermodynamic point of view (Table 5), reaction (34) is possible from a temperature of 200 °C. The above thermal decomposition can be combined with the thermal decomposition of HgO according to reaction (35) since the leaching of HgO is also thermodynamically possible (Table 5) according to reaction (36). The HgO decomposition reaction is thermodynamically possible from 600 °C Table 5. At this temperature, Hg is in the gaseous state and can be removed from the feedstock and subsequently recovered by condensation. At 600 °C, the thermal decomposition of Ag2O will also be shifted with sufficient efficiency to the right-hand side of the equation. Under this assumption, silver will not pass into solution during leaching with H2SO4 in the form of Ag2SO4 (Equation (37)), because it is thermodynamically impassable (Table 5). After the oxides have been converted to solution by leaching with H2SO4 and after subsequent thermal decomposition, the noble metals Au and Ag remain in the leach, which can be obtained by smelting at 1100 °C.

3.4. Cementation of Metals from Solution to Obtain Marketable Products

The next technological step is the extraction of copper from the pregnant solution by its process of cementation on iron to prepare Cu powder or (Fe-6 Cu) powder according to the chemical reaction (38). This follows from the galvanic series of metals since the electrode potential of the redox pair Fe2+/Fe0 is more negative compared to the electrode potential of the redox pair Cu2+/Cu0 according to relation (40).
E F e 2 + / F e 0 0 < E C u 2 + / C u 0 0             − 0.41   V < 0.34   V .  
To completely remove the copper from the solution, an overplus of the metal from the cementing agent is required. The extracted powder is contaminated with the cementing metal. The iron impurity in the powdered copper may be around 2%. Solution CuSO4 is used for refining Cu powder from Fe impurities obtained by the reaction according to Equation (40). Waste copper metal oxides (95% Cu2O and 5% CuO) are formed during the production of copper wire and rod at a temperature of 900 °C. The product of Equation (38) is precipitated Cu powder and CuSO4 solution. The CuSO4 solution would then be used in the process to refine the Cu powder from the iron.
For the refining of the powder from iron impurities, CuSO4 is used, which is obtained according to Equation (40) at a concentration of 12 g/L acidified to pH 1.8–2.5 at a liquid-to-solid phase L:S ratio of 5:1 and a temperature of 50 °C. Using such a solution in the refining process would increase the copper content of the powder up to 99.5–99.8% Cu, depending on the impurity content of the starting solution. Impurities in cemented copper can be further removed by washing with sulfuric acid without an oxidizing agent. This process leaches base metals, including residual iron, into the solution while leaving the cemented metallic copper intact. Only copper oxides on the surface of the metallic copper undergo leaching, while the metal itself remains unaffected [43,44].
The Fe-6 Cu (Fe 94%) system was a highly desirable product for powder metallurgy in the past. The Fe powder required for cementation, or even for the (Fe-6 Cu) system, can be prepared by electrolysis from a FeSO4 solution according to well-established production methods commonly used in practice. At the cathode, Fe powder is precipitated to be used as a cementing agent, and at the anode H2SO4 will be formed, which is recycled to leach the oxides. The procedure for the preparation of Fe-6 Cu powder at a laboratory scale is given in [42,45]. Considering that Cu2O can be formed during the conversion of sulfides to oxides and H2SO4 is formed at the anode, the preparation of the so-called precipitated Cu powder according to the reaction (27) is possible.
It is important to note that the mechanical recovery of precipitation Cu powder would be problematic because it would remain as a solid phase in the residue after leaching. The precipitation powder is of the highest quality after the electrolytic powder compared to the cementation powder, i.e., the reduction-obtained and thermally sputtered powder. Copper metal oxides, of which about 95% is Cu2O and 5% CuO, are formed in the production of copper wire and rod at a temperature of 900 °C. With the use of hydrometallurgical technology equipment, the preparation of precipitation powder from copper metal oxides would be possible. The procedure for the preparation of precipitation Cu powder on a laboratory scale is presented in the final technical report [46] as well as compared with work on copper recovery from industrial wastewater [47]. Copper deposit morphology which is modified by such factors as pH and copper concentration appears to be a crucial factor influencing overall recovery kinetics [48].
The advantage of the mentioned hydrometallurgical technology is the obtaining of noble metals, namely Ag and Au, which will be concentrated in practically 100% quantity in the residue after leaching. Another advantage is the possibility of preparing Cu or (Fe-6 Cu) powder by cementation, which has applications in powder metallurgy and the chemical industry. In addition, Fe powder as a cementing agent is prepared by electrolysis from its raw materials or can also be a saleable product. From an ecological point of view, it is significant that this hydrometallurgical method forms a closed technological cycle, i.e., there is no pollution of water by technological waste from production.
The second alternative treatment of sulfide concentrate is shown in Figure 5. This proposed technology (Figure 4) shows that instead of preparing Fe by electrolysis, it is possible to obtain Fe in the form of Fe(OH)3 from FeSO4 after Fe2+ → Fe3+ oxidation and hydrolysis, as presented by chemical reaction (41):
2 F e S O 4 + H 2 O 2 ( g ) + 4 H 2 O = 2 F e ( O H ) 3 + 2 H 2 S O 4 .
Fe2O3 can be prepared by thermal decomposition at a temperature of 500 °C according to the reaction (42).
2 F e ( O H ) 3 = F e 2 O 3 + 3 H 2 O                 Δ G 500 0 − 106.687   k J .
The Fe2O3 obtained is a marketable product. During the hydrolysis process according to chemical reaction (41), H2SO4 is formed in the electrolysis of Fe and is returned to the leaching process of oxides.
The advantage of this alternative is that hydrolysis is economically more advantageous than electrolysis. The disadvantage is that Fe powder for cementation must be added externally.
Since the undesirable phenomenon of sintering occurs in the pyrometallurgical method of producing Cu powder, it is not practically possible to achieve the separation of the two components with acceptable efficiency.
Based on these assumptions, it would be possible to combine the hydrometallurgical method with the pyrometallurgical method to prepare a compact copper powder. A brief description of the proposal is as follows:
▪
Prepare bricks by pressing from Cu powder that has been produced by the hydrometallurgical method using the cementation process, according to chemical reaction (38 and 40).
▪
Remelt the said bricks in a medium-frequency induction furnace under an inert atmosphere and subsequently carry out an oxidative refining process in this furnace using Cu2O to convert the polluting elements Sb, As, Hg, and Fe in the form of oxides into the solid phase, according to chemical reactions (22) to (25).
▪
Pour the melt into a settling resistance furnace to leach oxides of Sb, As, Hg, and Fe to the surface of the melt and remove these oxides in the form of slag. After the slag is removed, the refined copper is poured into a casting [49].
Another possible application of Fe2O3 is as a pigment, where the final product should contain more than 65.4% Fe2O3. A high-purity iron oxide (α-Fe2O3) with 99.6% purity can be obtained by roasting γ-FeOOH at 700 °C for 2 h, meeting national standards for pigment use [50].

4. Conclusions

Primary sources of antimony are rare strategic resources for the EU. Antimony is irreplaceable in modern IT technologies. Slovakia has antimony resources and in the past was engaged in its production. During the production of antimony, various by-products such as copper and iron sulfides with gold and silver content are formed.
Based on the available thermodynamic data, a series of theoretical calculations on sulfide concentrate dissolution are studied in this research. The research results of this work are presented as follows:
▪
In the pyrometallurgical processing of Cu(Ag,Au)Fe sulfide concentrate during roasting, the undesirable phenomenon of the sintering of the Cu and Fe components prevents sintering of both components from taking place with acceptable efficiency.
▪
In the hydrometallurgical processing method, the technological cycle is closed, which is more advantageous both from the economic and ecological points of view. The conversion of CuO and Cu2O oxides into solution using H2SO4 is a thermodynamically feasible solution at 20 °C. According to the Eh-pH diagrams, copper is calculated in a thermo-dynamic system, and Cu2+ ions are formed at a Cu concentration of 1 mol/dm3 in the range below pH 2.5. In the case of iron, much depends on the oxidation conditions of the system.
▪
The product of the hydrometallurgical method of copper production is Cu powder and composite powders of the Fe-Cu system.
In the case of hydrometallurgical processing of the concentrate, with direct recovery of metallic antimony and mercury and sale of Cu(Ag,Au),Fe as a solid residue after Sb and Hg extraction, a higher economic benefit can be expected. In any case, the most economically advantageous would be complete hydrometallurgical processing, i.e., also Cu(Ag,Au),Fe sulfide concentrate, with the possibility of valorizing Cu, Ag, and Au in metallic form. Another economic benefit would be the production of antimony oxide, which is a higher form of finalization, directly from the cathode metal, after electrolysis. During processing, the energy-intensive benefit from the technological operation of antimony refining is excluded.

Author Contributions

Conceptualization, D.O. and M.L.; methodology, F.M. and J.K.; software, J.K. and V.M.; validation, D.O., M.L. and F.M.; formal analysis, V.M. and J.P.; investigation, M.L.; resources, M.L.; data curation, J.K. and V.M.; writing—original draft preparation, F.M.; writing—review and editing, M.L.; visualization, M.L.; supervision, D.O.; project administration, D.O. and J.P.; funding acquisition, D.O. All authors have read and agreed to the published version of the manuscript.

Funding

This work was funded by The Ministry of Education, Science, Research and Sport of the Slovak Republic under grant number VEGA 1/0678/23 and grant number VEGA 1/0408/23.

Data Availability Statement

Data are contained within the article.

Acknowledgments

We thank the Slovak Railways for making available historical information on the former production of antimony in Slovakia. Thanks also go to the Institute of Geotechnics of the Slovak Academy of Sciences, v. in. i. to perform analyses on the sample.

Conflicts of Interest

The authors declare no conflicts of interest.

References

  1. Critical Raw Materials Resilience: Charting a Path Towards Greater Security and Sustainability. Available online: https://eur-lex.europa.eu/legal-content/EN/TXT/?uri=CELEX:52020DC0474 (accessed on 10 December 2024).
  2. Laubertová, M.; Šándorová, K. Mineral wealth of the Caspian Region states and their prospective cooperation with the European Union. Metall 2014, 68, 248–254. [Google Scholar]
  3. Hammarstrom, J.M.; Dicken, C.L.; Woodruff, L.G.; Andersen, A.K.; Brennan, S.T.; Day, W.C.; Drenth, B.J.; Foley, N.K.; Hall, S.; Hofstra, A.H.; et al. Focus Areas for Data Acquisition for Potential Domestic Resources of 13 Critical Minerals in the Conterminous United States and Puerto Rico—Antimony, Barite, Beryllium, Chromium, Fluorspar, Hafnium, Helium, Magnesium, Manganese, Potash, Uranium, Vanadiu; Geology, Energy & Minerals Science Center: Reston, VA, USA, 2022. [Google Scholar]
  4. Anderson, C.G. The metallurgy of antimony. Chem. Erde 2012, 72, 3–8. [Google Scholar] [CrossRef]
  5. Zhao, G.; Li, W.; Geng, Y.; Bleischwitz, R. Uncovering the features of global antimony resource trade network. Resour. Policy 2023, 85, 103815. [Google Scholar] [CrossRef]
  6. Kanellopoulos, C.; Sboras, S.; Voudouris, P.; Soukis, K.; Moritz, R. Antimony’s Significance as a Critical Metal: The Global Perspective and the Greek Deposits. Minerals 2024, 14, 121. [Google Scholar] [CrossRef]
  7. Yang, J.G.; Wu, Y.T. A hydrometallurgical process for the separation and recovery of antimony. Hydrometallurgy 2014, 143, 68–74. [Google Scholar] [CrossRef]
  8. Dembele, S.; Akcil, A.; Panda, S. Technological trends, emerging applications and metallurgical strategies in antimony recovery from stibnite. Miner. Eng. 2022, 175, 107304. [Google Scholar] [CrossRef]
  9. Dupont, D.; Arnout, S.; Jones, P.T.; Binnemans, K. Antimony Recovery from End-of-Life Products and Industrial Process Residues: A Critical Review. J. Sustain. Metall. 2016, 2, 79–103. [Google Scholar] [CrossRef]
  10. Segura-Salazar, J.; Brito-Parada, P.R. Stibnite froth flotation: A critical review. Miner. Eng. 2021, 163, 106713. [Google Scholar] [CrossRef]
  11. Ukasik, M.; Jergová, K.; Laubertova, M.; Havlik, T. Leaching of Tetrahedrite calcine using ozone. Acta Montan. Slovaca 2006, 12, 411–418. [Google Scholar]
  12. Raschman, P.; Sminčáková, E. Kinetics of leaching of stibnite by mixed Na 2S and NaOH solutions. Hydrometallurgy 2012, 113–114, 60–66. [Google Scholar] [CrossRef]
  13. Ubaldini, S.; Vegliò, F.; Fornari, P.; Abbruzzese, C. Process flow-sheet for gold and antimony recovery from stibnite. Hydrometallurgy 2000, 57, 187–199. [Google Scholar] [CrossRef]
  14. Blistan, P.; Kršák, B.; Blistanová, M.; Ferencz, V. The seabed-an important mineral resource of Slovakia in the future. Acta Montan. Slovaca 2015, 20, 334–341. [Google Scholar] [CrossRef]
  15. Bačo, P.; Bačová, Z.; Németh, Z.; Repčiak, M. Potential Occurrence of Selected, Mainly Critical Raw Materials at the Territory of the Slovak Republic in Respect to EU Countries Needs. Slovak Geol. Mag. 2015, 15, 87–120. [Google Scholar]
  16. Kaufmann, A.B.; Lazarov, M.; Weyer, S.; Števko, M.; Kiefer, S.; Majzlan, J. Changes in antimony isotopic composition as a tracer of hydrothermal fluid evolution at the Sb deposits in Pezinok (Slovakia). Miner. Depos. 2024, 59, 559–575. [Google Scholar] [CrossRef]
  17. Baláž, P. Review of Reserved Deposits of Metals in Slovakia. Slovak Geol. Mag. 2015, 15, 31–38. [Google Scholar]
  18. Laubertova, M.; Trpčevská, J.; Pirošková, J.; Kostadinov, J. Antimony as a critical raw material for the European Union. Metall 2017, 71, 287–291. [Google Scholar]
  19. Radková, A.B.; Jamieson, H.; Lalinská-Voleková, B.; Majzlan, J.; Števko, M.; Chovan, M. Mineralogical controls on antimony and arsenic mobility during tetrahedrite-tennantite weathering at historic mine sites Špania Dolina-Piesky and L’ubietová-Svätodušná, Slovakia. Am. Mineral. 2017, 102, 1091–1100. [Google Scholar] [CrossRef]
  20. Sejkora, J.; Števko, M.; Macek, I. Příspěvek k chemickému složení tetraedritu z Cu ložiska Piesky, rudní revír Špania Dolina, střední Slovensko. Bull. Mineral. Oddel. Nar. Muz. Praze 2013, 21, 89–103. [Google Scholar]
  21. Havlík, T. Hydrometallurgy Principles and Applications; Woodhead Publishing Limited: Cambridge, UK, 2008; ISBN 978-1-84569-407-4. [Google Scholar]
  22. Neiva Correia, M.J.; Carvalho, J.R.; Monhemius, A.J. Study of the autoclave leaching of a tetrahedrite concentrate. Miner. Eng. 1993, 6, 1117–1125. [Google Scholar] [CrossRef]
  23. Xu, B.; Zhong, H.; Jiang, T. Recovery of valuable metals from Gacun complex copper concentrate by two-stage countercurrent oxygen pressure acid leaching process. Miner. Eng. 2011, 24, 1082–1083. [Google Scholar] [CrossRef]
  24. Havlik, T.; Skrobian, M.; Dudas, D. Study of oxide acidic leaching of Tetrahedrite. Hut. List. 1991, 46, 76–80. [Google Scholar]
  25. Sudova, M.; Kanuchova, M.; Sisol, M.; Kozakova, L.; Marcin, M.; Holub, T. Possibilities for the Environmental Processing of Gold-Bearing Ores. Separations 2023, 10, 384. [Google Scholar] [CrossRef]
  26. Correia, M.J.; Carvalho, J.; Monhemius, J. Effect of tetrahedrite composition on its leaching behaviour in FeCl3-NaCl-HCl solutions. Miner. Eng. 2001, 14, 185–195. [Google Scholar] [CrossRef]
  27. Baláž, P.; Ficeriová, J.; Šepelák, V.; Kammel, R. Thiourea leaching of silver from mechanically activated tetrahedrite. Hydrometallurgy 1996, 43, 367–377. [Google Scholar] [CrossRef]
  28. Safarzadeh, M.S.; Miller, J.D. Acid bake-leach process for the treatment of arsenopyrite, tennantite, and tetrahedrite. Int. J. Miner. Process. 2013, 124, 128–131. [Google Scholar] [CrossRef]
  29. Awe, S.A.; Sandström, K. Selective leaching of arsenic and antimony from a tetrahedrite rich complex sulphide concentrate using alkaline sulphide solution. Miner. Eng. 2010, 23, 1227–1236. [Google Scholar] [CrossRef]
  30. Baláž, P.; Achimovičová, M.; Ficeriová, J.; Kammel, R.; Šepelák, V. Leaching of antimony and mercury from mechanically activated tetrahedrite Cu12Sb4S13. Hydrometallurgy 1998, 47, 297–307. [Google Scholar] [CrossRef]
  31. Ukasik, M.; Havlik, T. Effect of selected parameters on tetrahedrite leaching by ozone. Hydrometallurgy 2005, 77, 139–145. [Google Scholar] [CrossRef]
  32. Guo, P.; Wang, S.; Zhang, L. Selective removal of antimony from refractory gold ores by ultrasound. Hydrometallurgy 2019, 190, 105161. [Google Scholar] [CrossRef]
  33. Lovás, M.; Murová, I.; Mockovciaková, A.; Rowson, N.; Jakabský, Š. Intensification of magnetic separation and leaching of Cu-ores by microwave radiation. Sep. Purif. Technol. 2003, 31, 291–299. [Google Scholar] [CrossRef]
  34. Sudová, M.; Sisol, M.; Kanuchova, M.; Marcin, M.; Kurty, J. Environmentally Friendly Leaching of Antimony from Mining Residues Using Deep Eutectic Solvents: Optimization and Sustainable Extraction Strategies. Processes 2024, 12, 555. [Google Scholar] [CrossRef]
  35. Aghazadeh, S.; Abdollahi, H.; Gharabaghi, M.; Mirmohammadi, M. Selective leaching of antimony from tetrahedrite rich concentrate using alkaline sulfide solution with experimental design: Optimization and kinetic studies. J. Taiwan Inst. Chem. Eng. 2021, 119, 298–312. [Google Scholar] [CrossRef]
  36. Achimovicova, M.; Balaz, P.; Briancin, J. Vplyv mechanickej aktivácie na selektivitu kyslého lúhovania tetraedritu. Acta Montan. Slovaca 2005, 10, 151–155. [Google Scholar]
  37. Baláž, P.; Achimovičová, M. Selective leaching of antimony and arsenic from mechanically activated tetrahedrite, jamesonite and enargite. Int. J. Miner. Process. 2006, 81, 44–50. [Google Scholar] [CrossRef]
  38. Kutuk, S. Morphology, Crystal Structure and Thermal Properties of Nano-Sized Amorphous Colemanite Synthesis. Arab. J. Sci. Eng. 2024, 49, 11699–11716. [Google Scholar] [CrossRef]
  39. Sekula, F. Technológia spracovania tetraedritového koncentrátu z lokality baňa Mária v Rožňave. Acta Montan. Slovaca 2008, 13, 50–57. [Google Scholar]
  40. Sekula, F.; Baláž, P.; Jusko, F.; Molnar, F.; Jakabský, Š. Hydrometallurgical technology for processing tetrahedrite concentrates from the Mária mine in Rožňava. Acta Montan. Slovaca 1998, 3, 149–156. [Google Scholar]
  41. Roine, A. Outokumpu HSC Chemistry for Windows: Chemical Reaction and Equilibrium Software with Extensive Thermochemical Database. In User’s Guide Outokumpu HSC Chemistry® Windows; Outokumpu Research Oy: Helsinki, Finland, 2002. [Google Scholar]
  42. Peng, T.; Xu, L.; Luo, L. Quantitative Investigation of Roasting-magnetic Separation for Hematite Oolitic-ores: Mechanisms and Industrial Application. Open Chem. 2017, 15, 389–399. [Google Scholar] [CrossRef]
  43. Molnar, F. Production of Powder Metals. Vyroba Praskovych Kovov; ALFA: Bratislava, Slovakia, 1985. [Google Scholar]
  44. Oráč, D.; Klimko, J.; Klein, D.; Pirošková, J.; Liptai, P.; Vindt, T.; Miškufová, A. Hydrometallurgical recycling of copper anode furnace dust for a complete recovery of metal values. Metals 2022, 12, 36. [Google Scholar] [CrossRef]
  45. Oriňáková, R.; Kupková, M.; Dudrová, E.; Kabátová, M.; Šupicová, M. The role of coating in the cellular material preparation. Chem. Pap. 2004, 58. [Google Scholar]
  46. Habashi, F. Metal Powders by HydrometallurgicalRoute—A Review. Interceram 2010, 59, 287–297. [Google Scholar]
  47. Al-Saydeh, S.A.; El-Naas, M.H.; Zaidi, S.J. Copper removal from industrial wastewater: A comprehensive review. J. Ind. Eng. Chem. 2017, 56, 35–44. [Google Scholar] [CrossRef]
  48. Gros, F.; Baup, S.; Aurousseau, M. Copper cementation on zinc and iron mixtures: Part 2: Fluidized bed configuration. Hydrometallurgy 2011, 106, 119–126. [Google Scholar] [CrossRef]
  49. Ahmed, I.M.; El-Nadi, Y.A.; Daoud, J.A. Cementation of copper from spent copper-pickle sulfate solution by zinc ash. Hydrometallurgy 2011, 110, 62–66. [Google Scholar] [CrossRef]
  50. Gong, S.; Chen, J.; Rao, M.; Wen, B.; Xiao, Y.; Huang, L. Recovery of valuable metals in acid-soluble residue from NdFeB waste and the preparation of high-purity iron oxide for ferrites. J. Environ. Chem. Eng. 2024, 12, 114459. [Google Scholar] [CrossRef]
Processes 13 00842 g001
Figure 1. Flowsheet of the tetrahedrite concentrate treatment.
Figure 1. Flowsheet of the tetrahedrite concentrate treatment.
Processes 13 00842 g001
Processes 13 00842 g002
Figure 2. Flowsheet of the technological cycle of sulfide Cu(Ag,Au),Fe concentrate treatment by combined hydro and pyrometallurgical methods to obtain refined Cu and Ag, Au in anode sludge.
Figure 2. Flowsheet of the technological cycle of sulfide Cu(Ag,Au),Fe concentrate treatment by combined hydro and pyrometallurgical methods to obtain refined Cu and Ag, Au in anode sludge.
Processes 13 00842 g002
Processes 13 00842 g003
Figure 3. Eh-pH diagrams of Cu, Fe, Bi, and As in Me–S–H2O systems at 80 °C.
Figure 3. Eh-pH diagrams of Cu, Fe, Bi, and As in Me–S–H2O systems at 80 °C.
Processes 13 00842 g003
Processes 13 00842 g004
Figure 4. Flowsheet of a closed cycle of treatment of sulfidic Cu(Ag,Au),Fe concentrate by the hydrometallurgical method (alternative to preparation of Fe by electrolysis).
Figure 4. Flowsheet of a closed cycle of treatment of sulfidic Cu(Ag,Au),Fe concentrate by the hydrometallurgical method (alternative to preparation of Fe by electrolysis).
Processes 13 00842 g004
Processes 13 00842 g005
Figure 5. Flowsheet of the closed treatment of Cu(Ag,Au),Fe sulfide concentrate by the hydrometallurgical method (alternative preparation of Fe(OH)3 and (Fe2O3).
Figure 5. Flowsheet of the closed treatment of Cu(Ag,Au),Fe sulfide concentrate by the hydrometallurgical method (alternative preparation of Fe(OH)3 and (Fe2O3).
Processes 13 00842 g005
Table 1. Chemical analysis of the sample.
Table 1. Chemical analysis of the sample.
Elements(wt. %)(g/t)
CuSbAsHgBiFeSSiMgAlKL.O.I. 2AuAgPt
34.7 0.150.320.030.3121.426.54.880.611.76<LoD 18.61273171.26
Note: 1 LoD, limit of detection; 2 L.O.I., loss on ignition.
Table 2. The standard Gibbs energy change for reactions (4) to (21).
Table 2. The standard Gibbs energy change for reactions (4) to (21).
Chemical Reaction ΔG°T (kJ/mol) No.
20 °C 80 °C
C u S + 2 N a O H ( a q ) + 2 O 2 ( g ) = C u O + N a 2 S O 4 ( a q ) + H 2 O −742.9−723.5(4)
C u S + 2 N a O H ( a q ) + 1.75 O 2 ( g ) = 0.5 C u 2 O + N a 2 S O 4 ( a q ) + H 2 O −688.4−672.4(5)
C u 2 S + 2 N a O H ( a q ) + 2 O 2 ( g ) = C u 2 O + N a 2 S O 4 ( a q ) + H 2 O −733.0−713.4(6)
2 C u F e S 2 + 8 N a O H ( a q ) + 8.5 O 2 ( g ) = 2 C u O + F e 2 O 3 + 4 N a 2 S O 4 ( a q ) + 4 H 2 O −3302.2−3219.9(7)
3 C u F e S 2 + 12 N a O H ( a q ) + 12.5 O 2 ( g ) = 3 C u O + F e 3 O 4 + 6 N a 2 S O 4 ( a q ) + 6 H 2 O −4853.7−4734.4(8)
3 C u F e S 2 + 12 N a O H ( a q ) + 11.75 O 2 ( g ) = 1.5 C u 2 O + F e 3 O 4 + 6 N a 2 S O 4 ( a q ) + 6 H 2 O −4690.4−4580.9(9)
2 C u F e S 2 + 8 N a O H ( a q ) + 8 O 2 ( g ) = C u 2 O + F e 2 O 3 + 4 N a 2 S O 4 ( a q ) + 4 H 2 O −3193.3−3117.7(10)
2 F e S + 4 N a O H ( a q ) + 4.5 O 2 ( g ) = F e 2 O 3 + 2 N a 2 S O 4 ( a q ) + 2 H 2 O −1880.3−1836.4(11)
F e S + 2 N a O H ( a q ) + 2 O 2 ( g ) = F e O + N a 2 S O 4 ( a q ) + H 2 O −815.0−797.0(12)
3 F e S + 6 N a O H ( a q ) + 6.5 O 2 ( g ) = F e 3 O 4 + 3 N a 2 S O 4 ( a q ) + 3 H 2 O −2720.9−2659.1(13)
2 F e S 2 + 8 N a O H ( a q ) + 7.5 O 2 ( g ) = F e 2 O 3 + 4 N a 2 S O 4 ( a q ) + 4 H 2 O −3105.6−3039.3(14)
A s 2 S 3 + 6 N a O H ( a q ) + 6 O 2 ( g ) = A s 2 O 3 + 3 N a 2 S O 4 ( a q ) + 3 H 2 O −2499.2−2442.9(15)
A s 2 S 3 + 6 N a O H ( a q ) + 7 O 2 ( g ) = A s 2 O 5 + 3 N a 2 S O 4 ( a q ) + 3 H 2 O −2705.8−2636.4(16)
S b 2 S 3 + 6 N a O H ( a q ) + 6 O 2 ( g ) = S b 2 O 3 + 3 N a 2 S O 4 ( a q ) + 3 H 2 O −2438.9−2382.3(17)
S b 2 S 3 + 6 N a O H ( a q ) + 7 O 2 ( g ) = S b 2 O 5 + 3 N a 2 S O 4 ( a q ) + 3 H 2 O −2662.6−2593.2(18)
B i 2 S 3 + 6 N a O H ( a q ) + 6 O 2 ( g ) = B i 2 O 3 + 3 N a 2 S O 4 ( a q ) + 3 H 2 O −2370.8−2314.3(19)
H g S + 2 N a O H ( a q ) + 2 O 2 ( g ) = H g O + N a 2 S O 4 ( a q ) + H 2 O −677.7−658.9(20)
N a 2 S O 4 ( a q ) + B a O + H 2 O = B a S O 4 + 2 N a O H ( a q ) −158.1−156.9(21)
(aq)—“an aqueous solution”.
Table 3. The standard Gibbs energy change for reactions (22) to (25).
Table 3. The standard Gibbs energy change for reactions (22) to (25).
Chemical ReactionΔG°T (kJ/mol)No.
20 °C80 °C
2 S b + 5 C u 2 O = S b 2 O 5 + 10 C u −112.208−106.363(22)
2 A s + 5 C u 2 O = A s 2 O 5 + 10 C u −43.367−37.532(23)
2 B i + 3 C u 2 O = B i 2 O 3 + 6 C u −53.778−51.326(24)
2 F e + 3 C u 2 O = F e 2 O 3 + 6 C u −297.742−295.053(25)
Table 4. The standard Gibbs energy change for reactions (26) and (31).
Table 4. The standard Gibbs energy change for reactions (26) and (31).
Chemical ReactionΔG°T [kJ/mol]No.
20 °C80 °C
C u O + H 2 S O 4 ( a q ) = C u S O 4 ( a q ) + H 2 O −44.014−40.711(26)
C u 2 O + H 2 S O 4 ( a q ) = C u S O 4 ( a q ) + C u + H 2 O −24.330−20.048(27)
F e 2 O 3 + C u 2 O + 4 H 2 S O 4 = 2 C u S O 4 ( a q ) + 2 F e S O 4 ( a q ) + 4 H 2 O −89.086−100.389(28)
F e 3 O 4 + C u 2 O + 5 H 2 S O 4 = 2 C u S O 4 ( a q ) + 3 F e S O 4 ( a q ) + 4 H 2 O −134.740−154.269(29)
Fe2O3 + 3H2SO4 = Fe2(SO4)3 + 3H2O−165.216−162.496(30)
Fe3O4 + 4H2SO4 = FeSO4 + Fe2(SO4)3 + 4H2O−266.152 −261.885 (31)
Table 5. The standard Gibbs energy change for reactions (32) to (39).
Table 5. The standard Gibbs energy change for reactions (32) to (39).
Chemical ReactionΔG°T [kJ/mol]No.
20 °C80 °C200 °C600 °C 800 °C
2 A g + 0.5 O 2 ( g ) = A g 2 O −11.510−7.501---(32)
A g 2 O + H 2 S O 4 ( a q ) = A g 2 S O 4 ( a q ) + H 2 O −71.224−77.150---(33)
A g 2 O = 2 A g + 0.5 O 2 ( g ) --−0.399−24.851−35.884(34)
H g O = H g ( g ) + 0.5 O 2 ( g ) --39.753−1.699−21.577(35)
H g O + H 2 S O 4 ( i a ) = H g S O 4 ( i a ) + H 2 O −15.005−14.386−--(36)
2 A g + H 2 S O 4 ( a q ) = A g 2 S O 4 ( a q ) + H 2 ( g ) 155.224143.659---(37)
F e s + C u S O 4 ( a q ) = C u ( s ) + F e S O 4 ( a q ) −144.393−142.455---(38)
C u 2 O + H 2 S O 4 ( a q ) = C u 2 ( s ) + C u S O 4 ( a q ) + H 2 O −24.332−20.053---(39)
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Oráč, D.; Laubertová, M.; Molnár, F.; Klimko, J.; Marcinov, V.; Pirošková, J. Thermodynamic Study Proposal of Processing By-Product Containing Au, Ag, Cu and Fe Sulfides from Antimony Ore Treatment. Processes 2025, 13, 842. https://doi.org/10.3390/pr13030842

AMA Style

Oráč D, Laubertová M, Molnár F, Klimko J, Marcinov V, Pirošková J. Thermodynamic Study Proposal of Processing By-Product Containing Au, Ag, Cu and Fe Sulfides from Antimony Ore Treatment. Processes. 2025; 13(3):842. https://doi.org/10.3390/pr13030842

Chicago/Turabian Style

Oráč, Dušan, Martina Laubertová, František Molnár, Jakub Klimko, Vladimír Marcinov, and Jana Pirošková. 2025. "Thermodynamic Study Proposal of Processing By-Product Containing Au, Ag, Cu and Fe Sulfides from Antimony Ore Treatment" Processes 13, no. 3: 842. https://doi.org/10.3390/pr13030842

APA Style

Oráč, D., Laubertová, M., Molnár, F., Klimko, J., Marcinov, V., & Pirošková, J. (2025). Thermodynamic Study Proposal of Processing By-Product Containing Au, Ag, Cu and Fe Sulfides from Antimony Ore Treatment. Processes, 13(3), 842. https://doi.org/10.3390/pr13030842

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