Recovery of Cu and Fe from a Sphalerite Concentrate by the MnO₂–KI Leaching Oxidation System

Abstract

This study examined the leaching behavior of copper and iron from a sphalerite concentrate in sulfuric acid utilizing an ensemble MnO₂–KI oxidizing system. The temperature was shown to significantly influence the leaching kinetics, with the efficiency notably improving between 40 °C and 80 °C. The introduction of KI affected the balance between sulfur passivation and oxidant availability, facilitating increased leaching efficiencies. At 3 wt% KI, maximum recoveries of 82.1% Cu and 85.3% Fe were achieved, which indicates a notable decrease in surface passivation. Kinetic study analysis revealed low activation energies of 28.90 kJ mol⁻¹ for copper and 18.94 kJ mol⁻¹ for iron, indicating that both processes proceed readily at moderate temperature regimes. Despite being diffusion-controlled, the mechanisms of dissolution are different: iron leaching is more complicated, involving pyrite oxidation, sulfur layer formation, transformation to marcasite, and ultimately iron (III) release, whereas copper leaching involves direct interaction of chalcopyrite with the oxidants, similar to the behavior of sphalerite.

1. Introduction

A crucial industrial metal, copper finds use extensively in electronics, plumbing, telecommunications, and electrical wiring [1]. It is critical for contemporary technology and infrastructure because of its superior thermal and electrical conductivity as well as its resistance to corrosion. To fulfill rising worldwide demand and promote advances in technology, copper must be extracted efficiently from chalcopyrite (CuFeS₂), the main copper-bearing material. Chalcopyrite often occurs in a dispersed form alongside other sulfide minerals (such as pyrite FeS₂, galena PbS, and sphalerite ZnS) in polymetallic ore deposits [2,3]. These minerals are typically concentrated by flotation and then processed. However, as richer ores become depleted, hydrometallurgical processes are attracting increased attention for treating lower-grade and complex sulfide concentrates [2,4]. Metals can be extracted from such challenging concentrates using acidic leaching techniques without emitting harmful smelting gases (like SO₂). It is important to note that acid leaching of chalcopyrite and its flotation concentrate cannot be fully understood without considering the broader context of sulfide mineral extraction in acidic media. In this regard, the leaching behavior of sphalerite and pyrite should also be taken into account, as their mechanisms share many similarities with those of chalcopyrite [3,5]. Chalcopyrite concentrate is leached in sulfuric acid with the use of oxidizing agents in an ordinary hydrometallurgical process to solubilize copper for recovery [5,6,7]. Compared to roasting, these techniques can reduce air pollution by producing byproducts like sulfuric acid or sulfate salts and achieving high metal recoveries. However, due to its refractory nature, the rate of chalcopyrite leaching is extremely slow under ambient conditions. Thus, it requires the employment of potent oxidizing agents to promote its oxidative leaching [8,9]. Despite chalcopyrite’s abundancy among all other copper minerals, it is known to be the most stable copper mineral with its structurally face-centered tetragonal lattice [10]. Unlike many other copper sulfides, chalcopyrite tends to passivate during leaching by forming a solid product layer (e.g., sulfur or polysulfides) on its surface, which limits contact with the leaching solution and impedes further reaction [5,7,11]. This inherent inertness of chalcopyrite has made the development of enhanced leaching strategies a persistent challenge in hydrometallurgy.

Several oxidizing agents and techniques have been employed to improve chalcopyrite leaching in acidic media. Hydrogen peroxide (H₂O₂) is a strong oxidizer that can significantly accelerate the initial dissolution of sulfide minerals by oxidizing sulfide (S²⁻) to both sulfates and elemental sulfur [12,13]. In practice, the use of H₂O₂ alone can be limited by the rapid decomposition of H₂O₂ and the formation of a passive sulfur layer on the mineral surface [14]. Recent studies suggest that the formation, structure, and persistence of this sulfur layer are not solely determined by the type or strength of the oxidant, but rather by the rate at which the oxidant is supplied, as well as by the formation of hydrogen sulfide as a key intermediate that subsequently transforms into elemental sulfur [15,16,17]. To address this, various studies demonstrated that leaching chalcopyrite in H₂SO₄–H₂O₂ systems is greatly enhanced by additives that prevent formation of compact sulfur layers on the surface [13,18], for example, the addition of certain compounds such as ethylene glycol [19,20], iodine/chloride and iodide/ferric sulfate media [21,22], amino acids [23], etc. Also, certain studies were performed in order to prove that kinetics may be augmented with mechanical [24] or microwave assistance [25], as it has been shown that those techniques inhibit or disrupt the sulfur film, thereby maintaining an active reaction interface. In addition to hydrogen peroxide, other oxygen-based oxidizing agents such as ozone and molecular oxygen can also be effectively utilized in chalcopyrite leaching. These oxidants can operate under atmospheric conditions, but their efficiency is often enhanced under elevated pressure in systems such as autoclaves [16,17,26,27,28,29].

Iron (III) salts (such as FeCl₃, Fe₂(SO₄)₃, and Fe(NO₃)₃) are widely used oxidizing agents for various sulfide concentrates. In acidic sulfate media, ferric ions can oxidize the sulfide in chalcopyrite, converting sulfide sulfur to elemental sulfur while Fe³⁺ is reduced to Fe²⁺. This redox couple (Fe³⁺/Fe²⁺) is well-known, and it has been extensively studied in sulfide leaching. The main reason for such interest is the fact that ores rich in chalcopyrite are almost always followed by pyrite. Another reason comes from the biotechnological studies of sulfide leaching where it has been found that iron oxidizing bacteria indirectly leach sulfides (via the Fe³⁺/Fe²⁺ couple) [30]. Fe²⁺ generated during leaching can be re-oxidized to Fe³⁺ by external oxidants (e.g., oxygen) or via bio-oxidation by Acidithiobacillus sp. bacteria. This creates a cyclic process where a small amount of Fe³⁺ can cycle repeatedly, thus oxidizing more sulfides from chalcopyrite [31,32]. Ferric salt leaching, for instance, is very effective at dissolving chalcopyrite; however, chloride-based leaching is highly corrosive and poses materials-handling challenges while iron (III) ions are prone to form jarosite even on pH values above 2.5 [33].

Persulfate compounds (e.g., sodium persulfate, Na₂S₂O₈) have been explored as alternative oxidants due to their ability to generate sulfate radical species. Upon activation (e.g., by heat, light or sonication), persulfate yields sulfate radicals (▪SO₄⁻), which are highly reactive oxidizing agents capable of oxidizing disulfides and sulfur to higher oxidation states [34]. Free radicals such as hydroxyls and sulfates can oxidize chalcopyrite and other sulfides more aggressively than molecular oxidants, thereby enhancing the metal leaching rate. However, such strong oxidation agents can lead to a low leaching rate when used on poor raw materials such as ores, especially when there is a possibility to expend radicals on side reactions, needless to mention that they are also costly [34,35].

Industrially practiced nitrogen species catalyzed pressure leaching has also been researched and successfully applied for treatment of low-grade chalcopyrite concentrates. This highly oxidizing system also offers rapid kinetics [5].

While the above oxidizing agents and techniques can improve the leaching rate and/or leaching efficiency of chalcopyrite, they often come with significant drawbacks. Many require high reagent doses or stringent conditions, leading to increased operational costs and safety hazards. Some produce undesirable by-products (for example, chlorine gas in chloride systems or NO_x in nitrite and nitrate systems) that pose environmental challenges. Processing polymetallic concentrates containing multiple sulfide minerals and interactions during leaching can lead to severe passive action. For instance, the leaching of one sulfide (e.g., pyrite or sphalerite) may generate intermediate polysulfide or sulfate species that precipitate or form films on another mineral’s surface. The formation of insoluble elemental sulfur layers or precipitates like jarosite on chalcopyrite surfaces is a well-documented impediment that hinders further leaching [36]. Such a passivation phenomenon helps explain why many conventional oxidizing systems fail to achieve high copper recoveries from chalcopyrite when multiple sulfides are present. Additionally, certain catalytic approaches (e.g., adding trace metals) that work in idealized tests are not always practical. For example, the presence of silver within a sulfide matrix (such as silver incorporated in pyrite) can greatly accelerate chalcopyrite oxidation by the formation of Ag₂S, which helps electron transfer (through the Ag⁺/Ag₂S redox couple) [37]. Most of the concentrates found in genuine environments, however, lack enough silver to take advantage of this action; therefore, it is not viable to purposefully add silver salts to the leach. All these considerations underscore the need for alternative oxidation systems that can overcome chalcopyrite’s refractory behavior, especially for complex feeds where multiple sulfides and gangue minerals may interfere with the leaching process.

A growing body of research has demonstrated that iodine (typically introduced as iodide, which is oxidized in situ to iodine) can dramatically enhance the leaching of chalcopyrite under mild conditions. In conventional ferric sulfate media, adding a small amount of KI has been shown to increase copper extraction rates significantly—for example, one study reported a 47% higher Cu recovery at 50 °C when 0.5 g/L KI was added, along with a sharp reduction in the apparent activation energy of chalcopyrite dissolution. Iodine is a powerful oxidant that attacks chalcopyrite to produce elemental sulfur, and it participates in a regenerative redox cycle: iodide is continuously oxidized to iodine (by agents like Fe³⁺, MnO₂ or NO₂) [38], and iodine in turn oxidizes chalcopyrite (thus extracting Cu²⁺ and reverting to I⁻). Several investigations confirmed that this I⁻/I₂ shuttle can overcome or delay passivation layers on chalcopyrite, thereby sustaining faster leach kinetics even at ambient temperatures [38,39]. A notable application is the JX Iodine Process, in which ferric sulfate solutions with low-concentration iodide were irrigated over chalcopyrite ore in column tests—the presence of iodine markedly improved copper recoveries compared to ferric leaching alone, although the degree of improvement varied with ore mineralogy [38]. Likewise, patented iodine-leaching methods have demonstrated that chalcopyrite (and even refractory sulfidic minerals like enargite and pyrite) can be leached effectively at room temperature by employing an I⁻/I₂ cycle with continuous regeneration of the oxidant [40,41]. Overall, these studies have established that iodine/iodide can serve as an efficient recyclable oxidant, yielding higher copper dissolution rates than conventional oxidants and enabling leaching of chalcopyrite under relatively mild conditions. In contrast to previous studies on pure chalcopyrite, the present work applied a MnO₂–KI oxidation system to a complex sphalerite concentrate (containing chalcopyrite and pyrite) and systematically examined how varying the iodine dosage and temperature affects the leaching efficiency and kinetics for both copper and iron. Detailed kinetic and thermodynamic calculations were performed in order to follow the oxidative tests of the initial concentrate.

2. Materials and Methods

The base material used for all advanced leaching processes was the sphalerite concentrate, received from a polymetallic ore deposit flotation plant (“Rudnik”, Rudnik, Serbia). The collected sample was crushed and homogenized by a vibro mill (KHD Humboldt Wedag AG, Köln, Germany). During the leaching tests and the determination of the structural properties, the full range of granulometric classes was employed.

2.1. Structural Characterization

The initial concentrate and the obtained samples during oxidation leaching were collected and carefully characterized in order to determine the presence of the constituents and the surface morphology.

X-ray diffraction (XRD, Philips PW 1710/1820, Eindhoven, The Netherlands) was used to detect the phase arrangement by determining the mineral compositions of the concentrate and leaching residues. An automated diffractometer with a Cu tube running at 40 kV and 30 mA in the 2θ range of 4–65° was employed for gathering XRD patterns.

Scanning electron microscopy (SEM, JEOL JSM-7001F, Tokyo, Japan) equipped with an energy-dispersive spectrometer (EDS, Oxford Xplore 15, High Wycombe, UK) was used to depict surface functionalities and the mapping of sample surfaces.

In order to recognize metallic and non-metallic minerals, qualitative ore microscopic (OM) studies were conducted using a polarizing microscope (Carl Zeiss-Jena Axioscope 5 Pol, Jena, Germany) under reflected light. A color camera (Axiocam 105 color, White Plains, NY, USA) was connected to the microscope. With the “Multiphase” module of the Carl Zeiss AxioVision SE64 Rel. 4.9.1 software package, photomicrographs were gathered.

2.2. Leaching Experiments

Advanced leaching oxidation tests were performed according to our previous work (Figure 1) [42]. Leaching experiments were conducted in a closed glass reactor (Zhengzhou Great Wall, Xingyang, China). The reactor was adapted with a helix stirrer and temperature controller.

The reaction system consisted of base concentrate (10.0 g) and aqueous sulfuric acid solution (Zorka p. a, Šabac, Serbia), with a constant agitating rate of 300 rpm. Several operational variables were altered in the leaching tests, including temperature (40, 60, 70, and 80 °C), duration (0–180 min), and the presence of additional oxidizers–manganese dioxide (Alkaloid, purity > 95%, Skopje, North Macedonia) and potassium iodide 1–3 wt% of the initial concentrate (Merck, purity > 95%, Darmstadt, Germany). Manganese dioxide was dosed on a stoichiometric basis relative to the total sulfide-sulfur in the concentrate targeting complete oxidation of sulfide minerals to sulfate. Potassium iodide (KI), used as an auxiliary oxidant, was investigated up to a practical upper limit of 3 wt% of the initial concentrate mass to balance reactivity and cost. At determined time periods, acidic samples were collected and analyzed by atomic absorption spectrophotometry (AAS, Perkin Elmer 703, Shelton, CT, USA). To reduce the risk of measurement error, all leaching tests were conducted in triplicate, and the reported quantities of iron and copper are the average of the three measurements. The iron and copper extraction rates (%) were computed using AAS measurements.

The ratio of the chemical elements in the initial sample was determined by the AAS technique. The presence of the detected elements in descending order (Zn—46.91, S—28.87, Fe—6.77, Cu—0.806, Cd—0.338, Pb—0.287, As—1.8 × 10⁻⁴, all in wt%) was presented in our previous work [42]

The reaction solution was filtered to remove the leaching residues after each experiment was complete. After being gathered and thrice cleaned with distilled water, the solid residues were dried in an oven. Following preparation, the samples were carefully examined.

Kinetic models can be used to better describe various chemical reactions. Hence, we used the modified Sharps method of decreased half reaction time [43] to linearize the experimental results, utilizing various equations. Numerous articles provide a full description of the concept [43,44,45].

The Arrhenius equation, which is commonly used to calculate the activation energy (Ea), often provides important insights into the mechanisms driving the reactions:

k = A × exp(−Ea/RT)

(1)

where A is the pre-exponential constant, k is the rate constant (s⁻¹), Ea is the activation energy (J·mol⁻¹), R is the gas constant (8.314 J·mol⁻¹·K⁻¹), and T is the absolute temperature (K). The activation energy can be computed by plotting ln (k) versus 1/T, which provides the slope of the linear fit.

3. Results and Discussion

3.1. Material Characterization

To determine chemical, morphological and mineralogical properties, all collected specimens were thoroughly analyzed by XRD, AAS, SEM-EDS, and OM.

3.1.1. Optical Microscopy

From the OM results, nearly 80% of the sphalerite was found to exist as free grains (Figure 2a,b), with the remainder belonging to intergrowths with other minerals. In this concentrate, sphalerite is most commonly linked to chalcopyrite, which is followed by galena and chalcopyrite impregnations. The determined sphalerite’s grain size is in the range 10–400 μm. Gold and silver particles were also found in the observed concentrate (Figure 2b).

Chalcopyrite can appear as oval, micronic impregnations within sphalerite grains or as complex assemblages with galena and sphalerite minerals. It can also occur as individual grains, which makes up about 65–70% of its frequency. There was no detectable free gold or silver linked to chalcopyrite. A material specimen undergoing optical assessment contained a trace amount of gold or silver (Figure 2b).

3.1.2. XRD Analysis

The X-ray diffraction technique was used to analyze the initial polycrystalline concentrate. The analyzed sample was found to contain the following minerals: chalcopyrite, pyrite, quartz, and sphalerite [42]. Sphalerite is with certainty the most frequent mineral, while quartz, pyrite, and chalcopyrite are far less abundant.

3.1.3. Field Emission SEM

Initial concentrate particles possess clean and sharp edges, as seen in the SEM image (Figure 3). The particles have a diameter between 50 and 100 µm.

3.2. Leaching Assays

3.2.1. Advanced Oxidative Leaching of Initial Concentrate with MnO₂/KI

Impact of Temperature on Leaching Degree

The leaching efficiency of iron (Fe) and copper (Cu) from Zn concentrate increases with temperature, rising from 40 °C to 80 °C, in accordance with chemical kinetic principles. Elevated temperatures provide additional thermal energy, which enhances particle mobility and increases the frequency of the effective collisions between the sulfide minerals and the oxidizing agent (I₂). As a result, the reaction rate accelerates, leading to more efficient iron and copper extraction into the leaching solution.

The mineral phase distribution resulting from leaching at lower temperature is presented in Figure 4.

The analyzed sample (Figure 4) contains the following mineral phases: pyrolusite, sphalerite, chalcopyrite, marcasite, sulfur and galena, as well as gangue minerals (quartz, calcite, and barite). Pyrolusite is the dominant mineral, followed by sphalerite present in a lower proportion, while quartz, chalcopyrite, and marcasite are present in minor quantities. All other minerals are present in traces.

The quantitative mineralogical composition is as follows [42]: pyrolusite—54.3 wt%, sphalerite—41.2 wt%, chalcopyrite—1.1 wt%, marcasite—1.0 wt%, quartz—0.8 wt%, sulfur—0.6 wt%, calcite—0.4 wt%, galena—0.3 wt%, and barite—0.3 wt%. These values were determined using software processing, aligned with the results of chemical analysis.

The structural and textural characteristics, along with the mineral occurrence forms, are presented in Figure 5. Reflected-light microscopy analysis (Figure 5) indicates that pyrolusite is the predominant mineral, followed by sphalerite. Chalcopyrite, marcasite, and quartz are notably less abundant, while galena and sulfur appear only in trace amounts.

From the reflected light microscopy images (Figure 5) it can be observed that pyrolusite is predominantly found in a lath-like habitus, reaching lengths of up to 200 µm, and less commonly appears as irregular grains measuring up to 100 µm. Sphalerite typically occurs as irregular, liberated grains—often significantly altered and eroded—up to 150 µm in size. It frequently exhibits signs of “chalcopyrite disease” and contains chalcopyrite inclusions reaching up to 50 µm. Minor intergrowths of sphalerite with chalcopyrite, or marcasite are also observed. Apart from these occurrences, chalcopyrite is occasionally present as liberated, irregular grains up to 150 µm in length. Marcasite is found as free grains up to 150 µm in length.

Apart from the above, irregular galena grains up to 60 µm in length were also identified in the sample. The occurrence of galena in the sphalerite concentrate is anticipated, given that the original ore contains both sphalerite and galena, which are often challenging to separate due to their intergrowth—particularly when grain sizes are below 80 µm. Consequently, the presence of galena in the leaching residue is to be expected.

The SEM image of the leaching residue obtained at 40 °C is presented in Figure 6. The chemical compositions of the minerals obtained by EDS analyses are shown in

Table 1. EDS chemical analyses of spectra obtained from the SEM image (Figure 6) (Spectra 50–52).

Element	Spectrum 50 (Sphalerite)		Spectrum 51 (Pyrolusite)		Spectrum 52 (Chalcopyrite)
	Weight%	Atomic%	Weight%	Weight%	Weight%	Atomic%
O			30.83	60.19
Al			0.22	0.25
Si			0.60	0.67
S	33.38	49.81	0.28	0.27	35.18	50.28
Mn	0.42	0.37	66.89	38.04	0.41	0.34
Fe	12.05	10.33			29.69	24.36
Cu	0.36	0.27			33.02	23.82
Ni			0.23	0.12
Zn	53.30	39.02	0.95	0.45	1.71	1.20
Cd	0.48	0.20
Total	100.00	100.00	100.00	100.00	100.00	100.00

. As observed in Figure 6, the particles of the concentrate leached at a lower temperature exhibit signs of partial decomposition, evidenced by roughened and/or more rounded edges and surface alterations.

The mineral phase distribution resulting from leaching at higher temperature, 80 °C, is shown in Figure 7.

The analyzed sample (Figure 7) contains the following mineral phases: pyrolusite, sphalerite, chalcopyrite, marcasite, sulfur and galena, as well as gangue minerals (quartz, calcite and barite). Pyrolusite is the dominant mineral, followed by sphalerite in smaller amounts, while quartz, chalcopyrite, sulfur, and marcasite are present in minor quantities. All other minerals are present in traces.

The quantitative mineralogical composition [42] is as follows: pyrolusite—57 wt%, sphalerite—38 wt%, chalcopyrite—1.1 wt%, sulfur—1.0 wt%, marcasite—1.0 wt%, quartz—0.9 wt%, calcite—0.4 wt%, galena—0.3 wt%, and barite—0.3 wt%. These values were determined using software processing (Carl Zeiss AxioVision SE64 Rel. 4.9.1 with the “Multiphase” module), aligned with the results of chemical analysis.

The structural and textural characteristics, along with the mineral occurrence forms analysed by reflected-light microscopy, are presented in Figure 8.

Reflected-light microscopy observations (Figure 8) indicate that pyrolusite is the predominant mineral, followed by sphalerite. Chalcopyrite, marcasite, sulfur, and quartz are notably less abundant, and galena appears only in trace amounts. Pyrolusite primarily appears in a lath-like habitus, reaching lengths of up to 200 µm, with occasional irregular grains reaching 100 µm. Sphalerite typically occurs as irregular, eroded grains up to 250 µm and frequently exhibits “chalcopyrite disease” or contains chalcopyrite inclusions up to 50 µm in length. Less commonly, sphalerite and chalcopyrite are intergrown. Apart from these occurrences, chalcopyrite is occasionally present as liberated, irregular grains up to 100 µm. Marcasite is found as liberated grains up to 150 µm, while sulfur appears in the form of irregular grains. Gangue minerals mainly consist of quartz, with lesser amounts of feldspar, barite, and calcite.

The SEM image of the residue formed after leaching at 80 °C is shown in Figure 9, while the chemical composition of minerals obtained by EDS analyses is shown in

Table 2. EDS chemical analyses of spectra obtained from the SEM image (Figure 9) (Spectra 7–10).

Element	Spectrum 7 (Sphalerite)		Spectrum 8 (Chalcopyrite)		Spectrum 9 (Pyrolusite)		Spectrum 10 (Marcasite)
	Weight %	Atomic %	Weight %	Weight %	Weight %	Atomic %	Weight %	Atomic %
O					30.39	59.90
S	32.47	48.68	34.27	49.19	0.33	0.33	38.88	52.56
Mn	0.83	0.73	2.29	1.92	69.28	39.77	0.66	0.52
Fe	11.91	10.26	29.71	24.48			60.46	46.92
Cu	1.98	1.50	33.29	24.11
Zn	52.81	38.84	0.43	0.31
Total	100.00	100.00	100.00	100.00	100.00	100.00	100.00	100.00

.

As shown in the previously presented results, there are differences in the mineral content between the initial sample and those leached at higher and lower temperatures. In the initial sample, pyrite occurs as the primary iron-bearing mineral. The leaching treatments induced changes in the chemical environment, leading to oxidative conditions under which pyrite undergoes partial alteration. This process results in a decrease in pH and an increase in the concentration of the dissolved iron. Consequently, these conditions facilitate the transformation of pyrite into the less stable secondary phase, marcasite. Additionally, marcasite may be formed as a result of iron (Fe) leaching from sphalerite through a mechanism analogous to that observed in the case of pyrite. When conditions are favorable for the formation of iron sulfide with an orthorhombic FeS₂ structure, marcasite can precipitate from the solution. This process is typically associated with acidic environments and is often influenced by the presence of other reactive iron sulfide minerals, such as pyrite. Furthermore, during the leaching process, sphalerite undergoes decomposition, leading to the partial release of chalcopyrite, which occurs within the sphalerite matrix as inclusions or impregnations—a phenomenon known as “chalcopyrite disease”.

The occurrence and behavior of sulfur are discussed in detail in the study ref. [42].

The amount sof copper (Cu) and iron (Fe) leaching from the concentrate over the time are shown in Figure 10, with determined α (%).

The reaction kinetics and extraction efficacy were significantly impacted by rise in the temperature (Figure 10). After 3 h, the maximum recoveries for copper and iron were 82.1% and 85.3%, respectively. The formation of the passivation layers on the surface of chalcopyrite and marcasite grains may be the cause of the lower leaching percentages of copper and iron at lower temperatures [46,47].

With a sufficient MnO₂ dosage, the most probable mechanisms for copper and iron leaching from chalcopyrite and pyrite, respectively, are as follows:

2CuFeS₂ + 17MnO₂ + 18H₂SO₄ = 2CuSO₄ +Fe₂(SO₄)₃ + 17MnSO₄+ 18H₂O

(2)

2FeS₂ + 15MnO₂ + 14H₂SO₄ =Fe₂(SO₄)₃ + 15MnSO₄+ 14H₂O

(3)

Both reactions (Equations (2) and (3)) take place in the presence of potassium iodide and without it; however, there are side reactions that may impede or even stop these two reactions. Passivation layers that could form include sulfur, anglesite, and barite. The presence of iodine that readily forms in the chemical reaction (Equation (4)) impacts the erosion of the sulfur passivation layer and thus enhance further leaching of chalcopyrite.

MnO₂ + 2KI + 2H₂SO₄ = MnSO₄ + K₂SO₄ + I₂ + 2H₂O

(4)

The product of reaction (4) is iodine (I₂) which reacts further by capturing an iodide ion (I⁻) [48], and forms the triiodide anion as reported in the literature [49]. In the subsequent reaction (Equation (5)), iodine leaches the copper and is reduced back to iodide anion.

CuFeS₂ + 2I₂ = Cu²⁺ + Fe²⁺ + 2S + 4I⁻

(5)

While hydrometallurgical studies typically consider only reaction (5) for chalcopyrite oxidation by iodine, fundamental inorganic chemistry indicates that Cu²⁺ can also oxidize iodide ions to iodine, forming copper (I) iodide [50]. Although CuI is insoluble in water, its formation under such oxidative conditions is unlikely, as Cu⁺ is readily re-oxidized to the more stable Cu²⁺ state, rendering reaction (5) effectively the only viable pathway.

Besides the Cu²⁺ that is formed in reaction (5) another very important product is the Fe²⁺ ion. Ferrous ion is also unstable and in oxidative media it oxidizes to ferric ion, as reaction (6):

2Fe²⁺ + MnO₂ + 2H₂SO₄ = 2Fe³⁺ + SO₄²⁻ + MnSO₄ + 2H₂O

(6)

The ferric ion participates in several redox reactions and it can leach pyrite as well [51,52]. This is very important since the results of a recent study concluded that pyrite grains remain intact after 14 days of leaching chalcopyrite concentrate in the presence of iron (III) sulfate and potassium iodide [39]. The consecutive reactions proposed to account for iron (III) leaching from pyrite are presented in the subsequent reactions (Equations (7)–(9)):

FeS₂(pyrite) + xFe³⁺ = FeS^+x_2−x(pyrite) + xFe²⁺ + xS

(7)

FeS^+x_2−x(pyrite) + xH₃O⁺ = Fe_1−xS_2−2x(marcasite) + xHS⁻ + xFe³⁺ + xH₂O

(8)

H⁺ +HS⁻ + MnO₂ + H₂SO₄ = MnSO₄ + S + 2H₂O

(9)

Marcasite formed in reaction (8) continues to oxidize further in the same manner as pyrite in reaction (7) leading to progressive ferric leaching.

3.2.2. Influence of KI Amount on Leaching Effectiveness

The quantity of potassium iodide (KI) in the reaction solution at 80 °C shows a non-linear correlation with the leaching rate for copper (Cu) and iron (Fe) from sphalerite concentrates (Figure 11). This unusual trend points to an intricate interaction between the KI amount and the following chemical transformations that affects the oxidizing agents’ accessibility and efficacy during the observed leaching.

At a low KI content (1 wt% relative to the mass of sphalerite), the system operates under stable equilibrium conditions, where the iodine (I₂) formed in reaction (4) promotes more efficient oxidation in reaction (5), without the occurrence of harmful side-effects. At the above KI concentration, the degree of copper and iron leaching is high and amounts to 71.4% and 75.3%, respectively. This indicates that elemental sulfur, which was formed during the leaching reaction at a KI content of 1 wt%, does not significantly interfere with the leaching process of the above metals.

With an increase in the concentration of KI to 2 wt%, there is a unexpected reduction in the degree of leaching of copper (59.9%) and iron (63.1%). This can be attributed to the formation of elemental sulfur (S), which forms a passive layer on the surface of the mineral, preventing the further penetration of oxidizing agents and thereby hindering the leaching process of these metals.

When the KI concentration increases to 3 wt%, there is a sudden increase in the degree of leaching of copper (82.1%) and iron (85.3%). At higher iodine concentrations, the oxidant becomes sufficient not only to react with chalcopyrite but also to react with the elemental sulfur passivation layer formed in reactions (5), (7), and (9), converting it into soluble sulfur species with higher oxidation states. These include the formation of aqueous sulfur dioxide (SO_2(aq)), followed by the oxidation of thiosulfate to tetrathionate, which is ultimately oxidized further to sulfate ions. The oxidation of sulfur containing ion species was described in detail in our previous study [42].

When the initial potassium iodide concentration is higher, more iodine is produced which can break down the passivation layer, thus re-exposing the sphalerite surface to the acid and making the leaching process more efficient. In addition, at this level the KI system exhibits a higher redox potential, and due to the excess iodide ions, the equilibrium shifts towards the formation of free iodine. This provides a sufficient amount of active oxidizing agent to sustain the process, which overcomes the limitations observed at 2 wt% KI.

3.3. Leaching Kinetics

To model the leaching kinetics, various kinetic models encompassed within the modified Sharp’s method were employed to describe the system’s behavior, as depicted in Figure 12. Among these, the Jander equation [43] for three-dimensional diffusion through a product layer (depicted by navy blue triangle symbols on Figure 12) provided the best fit to the experimental data at 80 °C, yielding a coefficient of determination (R²) of 0.98. This high correlation indicates that the Jander model accurately describes the leaching process under the studied conditions. Consequently, this model was selected for the fitting of all leaching experimental data obtained at various temperatures.

The results suggest that the rate of the leaching reaction is limited by the diffusion of copper, iron, and sulfide species through a product layer—most likely composed of elemental sulfur formed on the surface of the sphalerite particles. This product layer acts as a barrier, impeding mass transfer and slowing down the leaching rate [53]. The formation of such a sulfur layer is consistent with the findings reported in the literature, where elemental sulfur deposition on sphalerite particles during leaching processes has been observed to significantly decrease the leaching yield of the target metals [54,55].

Employing the Jander formula, the fitted data collected from the advanced leaching of sphalerite concentrate with MnO₂ and KI throughout the temperature range of 40 to 80 °C are shown in Figure 13. The proposed model shows a high coefficient of determination with the experimental data over the whole temperature range (R² > 0.97 for Fe and R² > 0.98 for Cu), suggesting a strong agreement.

The estimated leaching rate constants employing the modified Sharp method are 0.0012 min⁻¹ and 0.0011 min⁻¹ at 80 °C, and 0.0005 min⁻¹ and 0.0003 min⁻¹ at 40 °C for Fe and Cu, respectively. The obtained results from Figure 13 indicate that raising the temperature considerably accelerates the oxidation rate, increasing it by more than twice at 80 °C.

The obtained curves from Figure 13 present a good correlation of the laboratory results and the Jander equation. Therefore, the proposed model well depicts three-dimensional diffusion through a product layer, indicating that species migration via a product layer is the rate-limiting phase. The passivation layer of elemental sulfur developed on the sphalerite concentrate particle surface could be the slowest and the limitation step. Each temperature causes this layer to develop, albeit its quantity and effect on the rate of leaching may fluctuate. The existence of sulfur residues is verified even at the lowest leaching temperature (40 °C), while the crystallization of sulfur is only verified at the highest leaching temperature (80 °C). Increased leaching levels and elevated temperatures promote crystallization of sulfur onto the grain surface. Hence, this phenomenon can enhance the passivation effect by generating a more coherent and impermeable layer [42]. By adding enough potassium iodide (>3 wt%), this problem can be lessened as the concentration of iodine in the leaching solution will rise. Since iodine efficiently transforms elemental sulfur toward soluble sulfur compounds, the elevated iodine level permits the oxidation of the produced sulfur passivation layer. Increased leaching rates persist even at higher temperatures because of this oxidation, which lowers the mass transfer barrier and improves the leaching efficiency; this has already been noted in available studies [56,57].

3.4. Determination of Activation Energy for Concentrate Leaching

The activation energy (Ea), determined by utilizing the linearized version of the Arrhenius equation, is the last part in the kinetic analyses. The findings from Figure 14 display a high linear association (R² > 0.95 for Cu recovery and R² > 0.98 for Fe recovery). Equation (10) was used for estimating the activation energy based on the slope of this linear function:

Ea = −R × Slope

(10)

As previously validated, the computed values of 28.90 and 18.96 kJ/mol for Cu and Fe extraction show that diffusion is the leaching reaction’s limiting phase [54,55,56].

The calculated activation energy for the Cu and Fe leaching indicates that the oxidation process has a low energy barrier (activation energy of 28.90 kJ mol⁻¹ for Cu). Given its low activation energy, copper leaching is a diffusion-controlled process, which is consistent with the value found by Kaplun et al. [58]. The activation energy of iron leaching is even lower (18.94 kJ mol⁻¹); such a low value is not obtained in any literature available on the study of pyrite leaching. However, such low values are obtained in the study of Asta and Acero for leaching marcasite in acidic media [59]. This observation is consistent with the absence of pyrite in the leaching residue, where only marcasite was detected with a sub-stoichiometric sulfur content (Table 2). Most likely Fe³⁺ ions interact with pyrite in the first step oxidizing the disulfides and the low pH value influences non-stoichiometric polymorph to be crystalized as marcasite. Once marcasite is formed, the strong oxidation agent (pyrolusite) assisted by iodine as a participant of the sulfur oxidation oxidize the marcasite, and iron (III) is leached further.

Such low activation energies suggest that leaching proceeds rapidly at moderate temperatures due to the low energy barrier of the oxidation process. This observation is consistent with the experimental results, where the Jander equation provided the best fit. The low activation-energy values further confirm that the leaching of both copper and iron is governed by a diffusion-controlled mechanism.

4. Conclusions

In our study we presented the Cu and Fe recovery from sphalerite concentrates in acidic solution, applying the auxiliary combination of oxidants—MnO₂ together with KI. The influence of temperature variation had a significant impact on Cu and Fe recovery; there were quicker reaction kinetics from 40 °C to 80 °C and significantly increased leaching efficiency.

The addition of KI resulted in a complex interaction between sulfur passivation and oxidant accessibility. Cu and Fe leaching achieved 82.1% and 85.3% at 3 wt% KI, thus a reduction of the passivation layer on the grain surface of the sphalerite concentrate.

Furthermore, low activation energies of 28.90 kJ mol⁻¹ for Cu and 18.94 kJ mol⁻¹ for Fe were found in the kinetic evaluation of the system MnO₂/KI+H₂SO₄. Although both copper and iron leaching kinetics are diffusion-controlled processes, the mechanism of their leaching differs significantly. Chalcopyrite interacts directly with both oxidation agents, thus copper leaching is similar to zinc leaching from sphalerite. Iron leaching is more complex and includes the interaction of pyrite with ferric ions, the formation of a sulfur passivation layer, and pyrite alteration to marcasite which eventually lead to the final decomposition of marcasite and iron (III) leaching.

The procedure is more energy-efficient and profitable because of the low activation energies, and this indicates that the leaching reactions continue readily at comparatively lower temperatures. Higher Cu and Fe recovery rates are made possible by the improved reaction kinetics, which also increase the efficacy in general.

In summary, the present investigation demonstrates that the temperature and the KI amount have a significant influence on Cu and Fe recovery from sphalerite, with the maximum Cu and Fe leaching rate achieved at 80 °C and 3 weight percent KI.