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Article

Recycling Potential of Copper-Bearing Waelz Slag via Oxidative Sulfuric Acid Leaching

by
Pavel Grudinsky
1,*,
Ekaterina Vasileva
1,2 and
Valery Dyubanov
1
1
I.P. Bardin Laboratory of Issues of Complex Ore Metallurgy, A.A. Baikov Institute of Metallurgy and Materials Science, Russian Academy of Science, 49 Leninsky Prosp., 119334 Moscow, Russia
2
College of New Materials, National University of Science & Technology “MISIS”, 4 Leninsky Prosp., 119049 Moscow, Russia
*
Author to whom correspondence should be addressed.
Metals 2025, 15(3), 330; https://doi.org/10.3390/met15030330
Submission received: 6 March 2025 / Revised: 14 March 2025 / Accepted: 17 March 2025 / Published: 18 March 2025
(This article belongs to the Special Issue Separation, Purification and Extraction of Metals from Primary and Secondary Resources)
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Versions Notes

Abstract

:
Copper-bearing Waelz slag (CBWS) is a solid by-product of the Waelz process, the disposal of which faces significant environmental challenges. In this study, oxidative sulfuric acid leaching was applied for the recovery of valuable elements from a CBWS sample containing 26.23% Fe, 0.82% Cu, and 0.81% Zn. Experimental leaching was conducted at temperature ranges, durations, and solid-to-liquid (S/L) ratios of 25–90 °C, 5–240 min, and 0.05–0.5 g/cm3, respectively. The consumption rates of H2SO4 and H2O2 ranged within 9.18–15.29 mmol/g and 0–7.35 mmol/g, which, at a 1:4:1 g/cm3/cm3 ratio, were equal to 225–375 g/dm3 H2SO4 and 0–250 g/dm3 H2O2, respectively. Various oxidants such as H2O2, MnO2, air, oxygen, and Fe3+ ions were tested in the leaching experiments. The optimal leaching conditions were proven to be a temperature of 70 °C, duration of 180 min, S/L ratio of 0.2 g/cm3, and consumption rate of 13.4 mmol H2SO4/g. These leaching conditions led to the recovery of 96.1% Fe, 87.0% Cu, and 86.9% Zn with the addition of 2.94 mmol H2O2/g and 95.2% Fe, 84.7% Cu, and 67.5% Zn with the addition of 0.095 g MnO2/g. These results suggest that metallic iron particles contained in a CBWS sample complicate copper dissolution.

Graphical Abstract

1. Introduction

The modern copper metallurgical industry faces numerous challenges, among which the decreasing availability of high-grade copper raw materials is prominent [1,2]. As reserves of high-grade ores become increasingly deficient, the industry is compelled to process low-grade ores [3], which necessitate more energy-intensive and expensive techniques. Concurrently, the global demand for copper is continuing to rise [4] due to its indispensable applications such as electrical systems, renewable energy technologies, etc. Besides primary copper resources, alternative raw materials such as industrial by-products and metallurgical wastes are attractive. Secondary resources, including copper smelter slag [5] and dust [6,7], zinc leach residues [8], mine tailings [9,10], spent catalysts [11], and e-waste [12], have been considered to be valuable materials for copper recovery. Depending on their origin and processing route, these wastes typically contain 0.2–40% Cu [13]. Given the copper content in these wastes, the recovery of copper from them is rational due to economic and environmental reasons [14]. Among these secondary resources, Waelz slag, also known as Waelz clinker, deserves attention due to its elevated contents of valuable elements, often including copper [15,16].
Waelz slag is a by-product generated by the Waelz process [17], a pyrometallurgical method widely used for recycling zinc-rich by-products, wastes, and ores. This process involves the reduction and volatilization of zinc and other volatile metals, resulting in a slag enriched with non-volatile components such as iron, calcium, silicon, and various heavy metals, including copper. Globally, approximately 1.75 million tons of Waelz slag are produced annually [18]. Different types of Waelz slag exist, depending on the feed materials and operational conditions used. The first type includes Waelz slag derived from processing zinc leach residues [19], copper smelter dust [20], and other copper-bearing raw materials; it frequently contains appreciable gold and silver percentages. Such a recycling practice is common in zinc plants [21]. This type of Waelz slag often has a high copper content [22]. The second type of Waelz slag, with a lower copper content, is produced during the treatment of steelmaking dust [23] and oxide zinc ores [24]. Both Waelz slag types have a substantial residual zinc content, which complicates their recycling in the iron and steel industries [15].
It should be noted that the storage and disposal of Waelz slag face significant environmental challenges due to the potential leaching of hazardous elements into soil and groundwater [16]. Various recycling approaches have been proposed for the treatment of Waelz slag, such as stabilization using the addition of specific agents to mitigate leaching [25], incorporation into construction materials [26,27,28,29], and the recovery of valuable metals [23,24,30]. However, while chemical stabilization and incorporation into construction materials are preferable for the second type of Waelz slag owing to its uncomplicated composition, copper-bearing Waelz slag with significant contents of heavy metals is more suitable for the recovery of valuable elements such as copper and silver. The recovery of valuable elements from copper-bearing Waelz slag has been studied using magnetic separation and flotation [31], reduction smelting along with copper slag [32], autoclave leaching in ammonia–ammonium sulfate medium [33], and sequential sulfuric acid and nitric acid leaching [34]. Leaching methods are advantageous due to their ability to selectively dissolve target metals, which ensures efficient and cost-effective metal recovery [35].
Historically, in the early 20th century, copper was extracted from ores with more than 2% Cu [36], while mine tailings with a substantially lower copper content were dumped. Nowadays, these old tailings are reprocessed due to improvements in beneficiation technologies [37] and their copper grade of up to 0.6% exceeding that of current ores [38], which is usually in the range of 0.2–0.4% Cu [39]. Following the same trend, Waelz slag and other similar waste materials can eventually become viable feedstock for metal recovery. In addition, the environmental impact of secondary copper recovery is less than that of copper recovery from ores [40]. Therefore, it is crucial to explore efficient processing methods for such materials.
In this study, we investigate copper-bearing Waelz slag (CBWS) processing based on the dissolution of copper, zinc, and iron during the oxidation acid leaching stage. The method is tested to assess its efficiency using sulfuric acid leaching at 25–90 °C with the addition of H2O2 or MnO2 as an oxidant. Furthermore, based on the recovery degrees of valuable elements, as well as the characterization of the solid residue, the feasibility and advantages of the proposed approach are discussed.

2. Materials and Methods

2.1. Raw Materials

A CBWS sample taken from “Chelyabinsk zinc plant” JSC (Chelyabinsk, Russia) was used in the study. The sample was prepared using ball milling performed by a Fritsch Pulverisette 7 premium line (Fritsch GmbH, Idar-Oberstein, Germany) device. Figure 1 shows the particle size distribution of the ground CBWS sample, which was measured using the laser diffraction method by a Bettersizer 2600 analyzer (Dandong Bettersize Instruments Ltd., Dandong, China).
As follows from the plot, the CBWS sample had approximately 95% particles less than 50 μm, which is reasonable for leaching experiments. Table 1 shows the elemental composition of the CBWS sample used in the experiments.
The contents of the major elements in the sample show that the CBWS sample was derived from electric arc furnace dust processing as a basic burden material for the plant Waelz kiln. However, elevated concentrations of metals such as Cu, As, Pb, Sb, and Cd suggest that copper smelter dust or zinc leach residue might have been other burden components. Therefore, the sample appeared to be a hybrid of two the types of Waelz slag mentioned in the introduction.
Figure 2 provides the XRD pattern of the CBWS sample. According to the XRD pattern and our previous study [15], the CBWS sample contained the following major minerals: Ca2MgSi2O7, Ca2Al2SiO7, CaMgSiO4, Ca3MgSi2O8, Mg2SiO4, MgO, α-Fe, FeO, γ-Fe2O3, α-FeOOH, C, CaS, Cu, and Cr. Iron consisted of 61% metallic, 14.2% ferrous (Fe2+), and 24.8% ferric (Fe3+) forms. Zinc was distributed between ZnO (54.4%), Zn-containing silicates (14%), ZnS (16.4%), and different low-soluble Zn phases, including ferrite (15.2%), whereas copper was basically in metallic form. Based on the elemental composition and above-mentioned mineralogical data, the compound composition of the CBWS sample was calculated so that it corresponded to the values in Table 1. Table 2 represents the derived compound composition, which was used for the calculation of acid consumption.

2.2. Experiments

Figure 3 visualizes the main experimental procedures used in the study. The leaching experiments were performed using HJ-4B or HJ-6B (Changzhou Surui Instrument Co., Ltd., Changzhou, China) magnetic stirrers with hot plates in conical glass flasks with volumes of 50 or 100 cm3. Sulfuric acid and the oxidant in the required amounts and concentrations were added into the flask with the ground CBWS sample in the range of 2–8 g to provide the conditions of the experiments, namely, sulfuric acid and oxidant agent consumption, as well as an initial solid-to-liquid (S/L) ratio. In most of the experiments, the oxidant agent was freshly prepared hydrogen peroxide solution. The sulfuric acid and hydrogen peroxide solutions were consumed in the ranges of 9.18–15.29 mmol/g and 0–7.35 mmol/g, respectively. The initial S/L ratio was in the range of 0.05–0.5 g/cm3. The initial H2SO4/H2O2 volume ratio was kept constant at 4.
In particular, the bulk of the experiments was carried out at an initial solid-to-liquid ratio of 0.2 g/cm3, where 4 g of the CBWS sample, 16 cm3 of a H2SO4 solution, and 4 cm3 of a H2O2 solution were used, which resulted in a final ratio of the CBWS sample, acid solution, and hydrogen peroxide of 1:4:1 g/cm3/cm3. As a case in point, in order to provide the consumption of 11.22 mmol of H2SO4/g and 2.94 mmol of H2O2/g at the ratio of 1:4:1 g/cm3/cm3, we added 16 cm3 of 275 g/dm3 H2SO4 and 4 cm3 of 100 g/dm3 H2O2 solution to 4 g of the CBWS sample.
When manganese dioxide was considered an oxidant, MnO2 analytical reagent of 85% purity was added in the range of 0–0.1489 g/g, which is equivalent to 0–10 wt.% Mn4+ of the CBWS sample. When Fe2(SO4)3∙9H2O was considered a source of ferric ions, its chemically pure reagent was added with the consumption rate of 0.06 g Fe3+/g. In some other experiments, air of 20.7 vol.% purity or oxygen of >99.7 vol.% purity was introduced into the stirring solution through a perforated silicone tube inserted through a flask neck. A dedicated gas cylinder containing air or oxygen served as the supply source to introduce gas into the flask. To ensure a stable and controlled gas flow, a variable-area flowmeter was installed within the system. The gas flow rate was 0.1 dm3/min.
The experimental procedure included the agitation of the pulp slurry using a fluoroplastic stirring bar of 27.5 mm in length and 7 mm in diameter. The temperature regulation system consisted of a thermocouple within a quartz sheath inserted into the pulp slurry. After keeping the pulp slurry at a specified temperature in the range of 25–90 °C for a duration in the range of 5–240 min, it was filtered using a vacuum pump with a suction funnel of 100 mm in diameter and a flask of 250 cm3 in volume. The residue collected on the filter was then washed with 0.01 M H2SO4 and dried at 90 °C for 120 min. The resulting solution was diluted in a volumetric flask of 200 or 250 cm3 using 0.01 M H2SO4.
The recovery degree of a target element was calculated as follows:
%ε = (Cs × Vs)/(MCDWS × %E/100) × 100
where %ε—the recovery degree of a target element, %; MCDWS—the initial mass of the CBWS sample, g; %E—the content of a target element in the CBWS sample, wt.%; Cs—the concentration of the element in the diluted solution, g/dm3; and Vs—the volume of the diluted solution, dm3.
A mother leach solution to recover iron and copper was obtained under the determined optimal leaching conditions. The used volume of the solution was 50 cm3. In the first stage, 100 g/dm3 H2O2 was added into the solution at 70 °C until achieving an absence of Fe2+ ions, which was controlled using K2Cr2O7 titration with a barium diphenylamine sulfonate indicator. Then, it was neutralized to pH = 4 using a 200 g/dm3 NaOH solution, so iron hydroxide was precipitated. The iron-containing residue was separated from the solution by vacuum filtration. In the second stage, the pH value was adjusted to 3 and a stoichiometric proportion of zinc metallic powder of −0.054 mm was added into the solution obtained in the first stage. After agitation on a magnetic stirrer in ambient conditions for 30 min, the solution was filtered by vacuum filtration and diluted in a volumetric flask.
Gibbs free energy changes in the discussed reactions were computed in the Reaction module of the HSC Chemistry 9.9 software (Outotec, Pori, Finland) [41]. This software provides a relevant thermodynamic database of numerous compounds and ions.

2.3. Analysis Methods

For an inductively coupled plasma atomic emission spectroscopic (ICP-AES) analysis, a combination of HNO3, HF, and H2SO4 was used to dissolve the initial CBWS sample, as well as the leaching residue obtained under the optimal conditions. If any insoluble residue remained, it was fused at 950 °C using Na2CO3 and H3BO3 and then completely dissolved by 1M HCl leaching. The dissolved samples were analyzed using a Varian Vista Pro (Varian Optical Spectroscopy Instr., Mulgrave, Australia) device. The iron content in the copper-bearing leachate solutions was measured using the redox titration method using KBH4 as a reductant of Fe3+ to Fe2+ and K2Cr2O7 as an oxidant of Fe2+ to Fe3+ according to the paper [42]. Other elements in the leachate solutions were analyzed by the ICP-AES device. The sulfur and carbon contents in the leaching residue were measured using a CS-230 analyzer (LECO Corporation, St. Joseph, MI, USA).
The mineralogical composition of the leaching residue was determined by X-ray diffraction analysis (XRD) using the diffractometer Tongda TDM-20 (Dandong Tongda Science & Technology., Ltd, Dandong, China) with Cu-Kα radiation and a scan rate of 0.05° 2θ/min. The microstructure and local composition of the solid residue were studied by scanning electron microscopy (SEM) using Vega 3SB (Tescan, Brno, Czech Republic) with an energy-dispersive X-ray (EDX) microanalysis detector.

3. Results

3.1. Determination of the Sulfuric Acid and Hydrogen Peroxide Consumption

The calculation of the required amount of sulfuric acid to dissolve the minerals of the CBWS sample was carried out using the compound composition (Table 2) and the following possible reactions [43,44,45]:
Fe + H2SO4 = FeSO4 + H2↑
FeO + H2SO4 = FeSO4 + H2O
γ-Fe2O3 + 3H2SO4 = Fe2(SO4)3 + 3H2O
FeOOH + 1.5H2SO4 = 0.5Fe2(SO4)3 + 2H2O
FeAs + 1.5H2SO4 + 2O2 = 0.5Fe2(SO4)3 + H3AsO4
Ca2MgSi2O7 + 3H2SO4 = MgSO4 + 2CaSO4 + 2SiO2 + 3H2O
CaMgSiO4 + 2H2SO4 = MgSO4 + CaSO4 + SiO2 + 2H2O
Ca3MgSi2O8 + 4H2SO4 = MgSO4 + 3CaSO4 + 2SiO2 + 4H2O
Ca2Al2SiO7 + 5H2SO4 = Al2(SO4)3 + 2CaSO4 + SiO2 + 5H2O
CaS + H2SO4 = CaSO4 + H2S↑
KAlSiO4 + 2H2SO4 = 0.5Al2(SO4)3 + 0.5K2SO4 + H2SiO3 + H2O
NaAlSiO4 + 2H2SO4 = 0.5Al2(SO4)3 + 0.5Na2SO4 + H2SiO3 + H2O
Mg2SiO4 + 2H2SO4 = 2MgSO4 + SiO2 + 2H2O
MgO + H2SO4 = MgSO4 + H2O
MnS + 4H2SO4 = MnSO4 + 4H2O + 4SO2↑
ZnFe2O4 + 4H2SO4 = ZnSO4 + Fe2(SO4)3 + 4H2O
ZnS + 4H2SO4 = ZnSO4 + 4H2O + 4SO2↑
Zn2SiO4 + 2H2SO4 = 2ZnSO4 + SiO2 + 2H2O
ZnO + H2SO4 = ZnSO4 + H2O
Pb2SiO4 + 2H2SO4 = 2PbSO4 + SiO2 + 2H2O
PbO + H2SO4 = PbSO4 + H2O
Ni + H2SO4 = NiSO4 + H2↑
Cu + H2SO4 + 0.5O2 = CuSO4 + H2O
Cr + H2SO4 = CrSO4 + H2↑
P2O5 + H2SO4 = 2HPO3 + SO3↑
Sb2O3 + 4H2SO4 = 2H[Sb(SO4)2] + 3H2O
BaO + H2SO4 = BaSO4 + H2O
CaTiO3 + H2SO4 = CaSO4 + TiO2 + H2O
According to the calculation, the required amount of H2SO4 was determined to be 1.292 g per 1 g of the CBWS sample. This consumption rate was rounded up to 1.3 g/g and then converted into molar units of measurement, obtaining the value of 13.26 mmol/g.
It is well-known that Cu dissolution in hydrometallurgical copper plants occurs mainly through a reaction with ferric iron [46], as follows:
Cu + 2Fe3+ = Cu2+ + 2Fe2+
Hence, to dissolve copper, we used H2O2 addition, which oxidizes iron according to the following reaction:
Fe2+ + 0.5H2O2 + H+ = Fe3+ + H2O
The amount of hydrogen peroxide according to Reaction (31) required to oxidize iron completely was found to be 0.799 per 1 g of CBWS sample. Taking into account the probable reactions of H2O2 with other components of the CDWS sample or the leaching solution, approximately 20% excess over the calculated value was accepted, obtaining 0.1 g/g or 2.94 mmol/g.
The calculation of the required amounts of acid and oxidant was verified experimentally. Figure 4 shows an influence of the acid and oxidant consumption on the recovery degrees of iron, copper, and zinc.
As follows from the presented dependences, the calculation of the required amounts of components appears to be substantially correct. The recovery degrees of Fe, Cu, and Zn rose when increasing the H2SO4 consumption rate from 9.18 to 13.26 mmol/g. When the acid consumption reached 13.26 mmol/g or more, there was no increase in the recovery degrees of iron and copper (Figure 4a). H2O2 consumption to dissolve the main part of copper was significantly lower than that shown in the calculations—if 0.735 mmol/g was added, 89% Cu was dissolved. However, this consumption rate is insufficient to convert iron into Fe3+ form, which is important for further iron precipitation stages (see Section 4). Therefore, taking into consideration the calculations, the experimental verification, and the need for further iron precipitation stages, we fixed the values of 13.26 mmol/g and 2.94 mmol/g as the consumption rates of H2SO4 and H2O2, respectively, for subsequent experiments.

3.2. Experimental Optimization of Leaching Conditions

The best leaching conditions were determined by successive optimization of the temperature, S/L ratio, and duration under the fixed conditions. Figure 5 illustrates the influence of leaching temperature on the recovery degrees of Fe, Cu, and Zn.
As is clear from the plot, the best temperatures for sulfuric acid leaching lie in the range of 50–90 °C, where the recovery degrees of the elements were over 80%. Reduced Cu recovery degrees below 80% were at lower temperatures due to a low dissolution rate of copper. The best temperature was considered to be 70 °C, where the combined copper and zinc recovery degrees were slightly higher. Figure 6 depicts the effect of the S/L ratio at 70 °C on the recovery degrees of the metals.
As reflected by the recovery degree dependences, a higher S/L ratio led to a gradual reduction in the extraction degrees of iron and zinc. Increasing the S/L ratio up to 0.33 g/cm3 resulted in a drastic drop in the copper recovery degree from above 80% to about 5%. Therefore, an S/L ratio of more than 0.33 is unsuitable for CBWS leaching. The highest Cu and Zn recovery degrees were obtained at 0.05 g/cm3. With S/L ratios at 0.1 and 0.2 g/cm3, they were only about 10% lower. It is evident that a larger volume of the solution at 0.05 g/cm3 considerably increased the operational costs of industrial leaching, while an insufficient increase in the recovery degrees occurred. In addition, such low S/L ratios can dissolve calcium sulfate more readily, increasing the calcium content in the leaching solution, whereas it is preferable for calcium to remain in the residue. Therefore, 0.2 g/cm3 was identified as the most suitable S/L ratio, ensuring both efficient metal recovery and cost-effectiveness. Figure 7 shows the duration influence at 70 °C and an S/L ratio of 0.2 g/cm3 on the element recovery degrees.
According to the plot, the metal recovery degrees already achieved a plateau after 15–30 min at 70–80% for Cu and Zn and 95–99% for Fe. After a sloping rise from 120 min, the highest recovery degrees of Fe, Cu, and Zn were achieved at 180 min of the leaching and maintained approximately the same level up to 240 min. Therefore, 180 min was chosen as the optimal leaching time.
Thus, the optimal conditions for oxidative sulfuric acid leaching with 96.1% Fe, 87.0% Cu, and 86.9% Zn recovery degrees were consumption rates of 13.4 mmol H2SO4/g and 2.94 mmol H2O2/g, an S/L ratio of 0.2 g/cm3, a temperature of 70 °C, and a leaching duration of 180 min. Table 3 lists the composition of the diluted leaching solution and the recovery degrees of other important elements under these conditions.
The recovery degrees of elements in the sulfuric acid solution varied significantly. Along with Fe, the highest recovery degrees above 90% were for Mg, Mn, Cd, and P, indicating their strong leachability. In contrast, Pb and Ba had predictably minimal recovery degrees due to the low solubility of the corresponding sulfates [47] generated according to Reactions (21), (22), and (28). As suggested in accordance with Reactions (7)–(10), (14), (19), and (21), silicon remained almost undissolved. Chromium, nickel, and antimony showed moderate dissolution degrees. Calcium was most likely partly passed into the solution due to a certain solubility of calcium sulfate during the washing of the residue with water.

3.3. Characterization of the Leaching Residue

In order to elucidate the behavior of valuable elements during the leaching and consider the recycling potential of the generated residue under the optimal conditions, we characterized its elemental and phase compositions. Table 4 and Figure 8 show the elemental and phase compositions of the residue.
As follows from the presented data, the leaching residue primarily consists of Ca and S, which form calcium sulfate phases such as CaSO4 and CaSO4·2H2O, along with a significant amount of C from graphite and Si from amorphous silica. An amorphous ring with a diffraction angle of up to 30° in the XRD pattern (Figure 8) confirms the presence of SiO2. The application of this residue as a mineralizer in Portland cement clinker production is possible. The addition of CaSO4-based mineralizers during treatment leads to an interaction with alkali metals to form alkali sulfates, which can stimulate alite formation [48]. Furthermore, it is essential to note that amorphous silica and graphite can be favorable due to their potential to partly substitute silica feed component and fuel, respectively. However, although the majority of trace heavy metals are at relatively low concentrations, the Cr content in the residue is challenging for its recycling in Portland cement production [49]. Leaching tests for heavy metals to verify compliance with environmental regulations are a crucial factor in the recycling potential of the residue in the construction industry [50]. The elemental and phase compositions of the residue substantially resemble the composition of flue gas desulfurization gypsum [51], so the applied utilization and treatment methods can be the same [52,53].
Figure 9 presents a SEM microphotograph and EDX mapping images of the residue that reveals its minor phases. As can be seen, the sample contains undissolved metallic iron particles (Figure 9g); it is evident that, after such a long leaching duration, only initially large metallic grains can remain undissolved. Probably, based on the microstructure of the CBWS sample, a part of undissolved copper can be present in such particles [15]. Mg, Cr, and Al are found in the same particles, which are probably hardly soluble stable spinel and chrome spinel structures.
For a more detailed determination of undissolved copper, zinc, and iron phases, an SEM–EDX investigation of points in the residue was carried out. Figure 10 demonstrates the phases found, while Table 5 lists their elemental compositions.
These phases were proven to be metal sulfides, which are poorly soluble even under oxidative sulfuric acid leaching conditions [54,55,56]. As follows from Figure 6, the solubility of zinc sulfide can probably be increased using an S/L ratio of 0.05 g/cm3. It is well-known that iron–copper sulfides are refractory [57]. However, taking into account Figure 6, where a recovery degree of 98.7% Cu was achieved at an S/L ratio of 0.05 g/cm3, the iron–copper sulfide amount was insignificant. In general, the presence of such sulfides complicates the complete extraction of valuable elements.

3.4. Possibility of Oxidant Substitution

It is reported [58] that the application of hydrogen peroxide in large-scale leaching processes is doubtful due to its instability and high cost. In order to find more suitable oxidants, we tested air and oxygen blowing, as well as the addition of MnO2 and Fe3+ ions under the optimal leaching conditions obtained in Section 3.2. Figure 11 depicts a comparison of oxidant types’ effects on the recovery degrees of iron, copper, and zinc.
As can be seen from the bar chart, the oxidant type significantly influences the leaching efficiency of metals. The copper recovery degree is negligible if the CBWS sample is leached without an oxidant, as well as with air or oxygen blowing. In contrast, Cu is mainly dissolved with the addition of H2O2, MnO2, and Fe3+. The use of air and oxygen blowing slightly decreases the Fe and Zn recovery degrees; the decreasing Fe recovery is likely due to its more substantial evaporation during blowing. The best results are related to H2O2 addition. MnO2 addition results in high recovery degrees of Fe and Cu, while the Zn recovery degree is about 20% lower than the best results for H2O2. The addition of ferric ions shows moderate Zn and lower Cu recoveries than the H2O2 and MnO2 additions.
The data presented suggest that MnO2 is as effective as hydrogen peroxide as an oxidant for copper dissolution, while other oxidants have proven to be less effective. Similar to hydrogen peroxide, manganese dioxide oxidizes iron to form ferric ions according to the following interaction:
Fe2+ + 0.5MnO2 + 2H+ = Fe3+ + 0.5Mn2+ + H2O
Therefore, the mechanism of the H2O2 and MnO2 oxidizing effect is quite analogous. Nevertheless, it should be noted that the acid amount required according to Reaction (32) is twice as much compared with Reaction (31). On the one hand, decreasing the acidity of the leaching solution is likely to lead to a Zn recovery degree reduction; on the other hand, manganese dioxide can be more effective in neutralizing the solution for its subsequent purification, which is the reason for MnO2 application in industrial practice [59]. Taking into account the results of the oxidant comparison, we studied the influence of MnO2 consumption on the metal recovery degrees, as illustrated in Figure 12.
As it appears from the plot, regardless of MnO2 consumption, the Fe recovery degree remains consistently close to 95–96%. If no MnO2 is added, the Zn recovery degree is above 80%; it drops to approximately 63–69% after the addition of any MnO2 amount, which is likely due to the decreasing dissolution of the zinc sulfide and ferrite of the CBWS sample (Section 2.1). The Cu recovery degree is more sensitive to MnO2 addition. It sharply increases at the consumption of 0.04 g Mn4+/g from near zero to 74.3%, then grows higher up to 84.7%, achieving a plateau at 0.06 g Mn4+/g, which is equal to 0.095 g MnO2/g.
The results show that, besides hydrogen peroxide, MnO2 can be used as an efficient oxidant for CBWS leaching with the recovery of 95.2% Fe, 84.7% Cu, and 67.5% Zn at consumption rates of 13.26 mmol H2SO4/g and 0.095 g MnO2/g, an S/L ratio of 0.2 g/cm3, a temperature of 70 °C, and a duration of 180 min.

4. Discussion

The results of experimental oxidative sulfuric acid leaching showed quite high recovery degrees of copper and zinc, increasing the recycling potential of the CBWS sample. It is encouraging to note that this process can be integrated into existing zinc plant hydrometallurgical operations, where leaching is followed by iron precipitation and copper cementation [60]. The reagents used, including sulfuric acid and manganese dioxide, are typically available in zinc production. Therefore, the leach solutions obtained from CBWS processing can be treated to extract iron, copper, and other elements so that the purified solution meets the requirements for zinc electrowinning. In order to substantiate this point, we conveniently purified a mother leach solution containing 31,680 mg/dm3 Fe and 1040 mg/dm3 Cu obtained from CBWS oxidative leaching. As a result, iron precipitation using the neutralization of the solution up to pH = 4 followed by copper precipitation using zinc powder treatment at pH = 3 led to residual metal concentrations of 1.9 mg/dm3 Fe and 15.7 mg/dm3 Cu and the recovery of 99.96% Fe and 91.1% Cu. Methods for subsequently purifying such solutions to produce electrolyte-grade zinc solutions are well-established [59].
The critical factor of the discussed process is economic efficiency. Undoubtedly, compared to the feed used in zinc plants, CBWS contains significantly lower contents of zinc and other valuable metals. Moreover, the use of quite expensive oxidants such as hydrogen peroxide and manganese dioxide contributes to an increase of processing costs.
As follows from Figure 11, the application of air or oxygen blowing during leaching, which could be an attractive low-cost solution, is ineffective for copper recovery from the CBWS sample, although it is well-known practically that the dissolution of copper is possible according to the following reaction [50]:
Cu + 0.5O2(g) + 2H+ = Cu2+ + H2O
At the same time, the CBWS sample contains metallic iron, which can react with copper and ferric ions according to the following reactions:
Fe + Cu2+ = Fe2+ + Cu
Fe + 2Fe3+ = 3Fe2+
Table 6 and Figure 13 show the influence of temperature on the Gibbs energy change in Reactions (30)–(35) providing insight into the behavior of copper in the presence of metallic iron.
As can be seen from the plot, all the reactions presented are possible. If Reaction (33) proceeds, then the generated Cu2+ ions can be reduced to metallic copper again according to Reaction (34). Hence, the dissolution of metallic iron is a prerequisite for copper dissolution. In our case, the CBWS sample contains large metallic iron particles with inclusions of metallic copper [15], so these particles slowly dissolve during leaching, preventing copper dissolution even in the presence of air or oxygen blowing. Thus, in order to apply air or oxygen as the oxidant for copper, it is important to provide conditions for the dissolution of metallic iron. This can be achieved by obtaining Waelz slag with small iron particles, which can be ensured by the operating conditions of the Waelz process. Another way is to prevent the presence of metallic iron in Waelz slag using iron oxidation at the output of the Waelz kiln [61], which is considered as a technique to increase the efficiency of heat management.
As mentioned above, the application of H2O2 and MnO2 contributes to the oxidation of ferrous into ferric iron according to Reactions (31) and (32), respectively, followed by copper dissolution through Reaction (30). It should be noted that the presence of generated ferric ions facilitates the dissolution of metallic iron via Reaction (35). Until the main parts of metallic particles pass into the solution, Reaction (30) hardly occurs due to its less negative free Gibbs energy change compared with Reaction (35) (Figure 13). Therefore, the metallic iron in the CBWS sample in the case of H2O2 and MnO2 use also hinders copper dissolution. Probably, the presence of undissolved metallic iron in the leaching residue (Figure 9g) slightly reduces the copper recovery degree, which can be higher, as is shown by leaching at an S/L ratio of 0.05 g/cm3 (Figure 6).
Based on the results obtained (Figure 11), Fe3+ ion addition can be considered as a promising agent for copper dissolution from CBWS samples in future studies. It can be introduced into leaching feedstock in the form of various compounds or solutions containing ferric iron. As discussed earlier, ferric ions enhance the dissolution of metallic iron and directly oxidize copper through Reactions (35) and (30), respectively. This approach offers a more affordable alternative to expensive oxidants. The utilization of the residue in construction or other industries could also serve as an important strategy to improve process efficiency and should be prioritized in future research.
The primary advantage of the Waelz process is its adaptability to various feedstocks [17]. However, the integrated treatment of raw materials with diverse origins and compositions in the Waelz kiln can be unfavorable due to the generation of complex by-products with complicated recycling potential due to an elevated level of harmful impurities. In our case, the Waelz slag contains both oxide and sulfide components that reduce the metal recovery efficiency, complicating the recycling of the obtained products. Obviously, the treatment of mixed oxide–sulfide raw materials is challenging [62]. A key strategy is adjusting the feed and technology to derive Waelz slag suited for further processing, regardless of the specific recycling approach adopted.

5. Conclusions

As a result of this study, oxidative sulfuric acid leaching has been discovered as a suitable method to recover valuable elements such as copper, zinc, and iron from CBWS samples. The most effective oxidants used during the leaching were found to be H2O2 and MnO2. The data obtained suggest that the large metallic iron particles contained in CBWS complicate copper recovery and inhibit applying air or oxygen blowing to dissolve copper. In order to improve the selectivity and efficiency of oxidative leaching process, it is necessary to provide conditions for the rapid dissolution of metallic iron, which can be achieved by controlling the size of the metallic particles in the original CBWS sample.
The best leaching conditions were an S/L ratio of 0.2 g/cm3 with a consumption rate of 13.4 mmol H2SO4/g of the CBWS sample, which is equal to a 325 g/dm3 H2SO4 concentration, as well as a leaching temperature and duration of 70 °C and 180 min, respectively. Under these conditions, 96.1% Fe, 87.0% Cu, and 86.9% Zn were recovered into the solution using a consumption rate of 2.94 mmol H2O2/g, while the recovery degrees were 95.2% Fe, 84.7% Cu, and 67.5% Zn at a consumption rate of 0.095 g MnO2/g.

Author Contributions

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

Funding

This research was funded by the Russian Science Foundation, grant number 24-23-00507.

Data Availability Statement

The data presented in this study are available on request from the corresponding author due to privacy reasons.

Acknowledgments

The authors appreciate the Chemical Analytical Laboratory of the JSC “Design & Survey and Research & Development Institute of Industrial Technology” for chemical analysis.

Conflicts of Interest

The authors declare no conflicts of interest.

Abbreviations

The following abbreviations are used in this manuscript:
CBWSCopper-bearing Waelz slag
S/L ratioSolid-to-liquid ratio
XRDX-ray diffraction
SEMScanning electron microscopy
EDX detectorEnergy-dispersive X-ray detector

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Metals 15 00330 g001
Figure 1. Particle size distribution of the ground CBWS sample.
Figure 1. Particle size distribution of the ground CBWS sample.
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Figure 2. The XRD pattern of the CBWS sample.
Figure 2. The XRD pattern of the CBWS sample.
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Figure 3. A schematic diagram of the experimental procedure.
Figure 3. A schematic diagram of the experimental procedure.
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Figure 4. The effect of H2SO4 consumption with the addition of 2.94 mmol H2O2/g (a), as well as the effect of H2O2 consumption at the rate of 13.26 mmol H2SO4/g (b) on the recovery degrees of Fe, Cu, and Zn from the CBWS sample. The following conditions were fixed: S/L ratio of 0.2 g/cm3, temperature of 60 °C, and duration of 180 min. Top horizontal axis demonstrates H2SO4 and H2O2 concentrations used to provide corresponding consumptions at the fixed 1:4:1 g/cm3/cm3 solid–acid–oxidant ratio. Dotted lines indicate chosen optimal consumption rates.
Figure 4. The effect of H2SO4 consumption with the addition of 2.94 mmol H2O2/g (a), as well as the effect of H2O2 consumption at the rate of 13.26 mmol H2SO4/g (b) on the recovery degrees of Fe, Cu, and Zn from the CBWS sample. The following conditions were fixed: S/L ratio of 0.2 g/cm3, temperature of 60 °C, and duration of 180 min. Top horizontal axis demonstrates H2SO4 and H2O2 concentrations used to provide corresponding consumptions at the fixed 1:4:1 g/cm3/cm3 solid–acid–oxidant ratio. Dotted lines indicate chosen optimal consumption rates.
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Figure 5. The effect of leaching temperature on the recovery degrees of Fe, Cu, and Zn from the CBWS sample at the consumption rates of 13.26 mmol H2SO4/g and 2.94 mmol H2O2/g, S/L ratio of 0.2 g/cm3, and duration of 180 min.
Figure 5. The effect of leaching temperature on the recovery degrees of Fe, Cu, and Zn from the CBWS sample at the consumption rates of 13.26 mmol H2SO4/g and 2.94 mmol H2O2/g, S/L ratio of 0.2 g/cm3, and duration of 180 min.
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Figure 6. The effect of S/L ratio on the recovery degrees of Fe, Cu, and Zn from the CBWS sample at the consumption rates of 13.26 mmol H2SO4/g and 2.94 mmol H2O2/g, temperature of 70 °C, and duration of 180 min.
Figure 6. The effect of S/L ratio on the recovery degrees of Fe, Cu, and Zn from the CBWS sample at the consumption rates of 13.26 mmol H2SO4/g and 2.94 mmol H2O2/g, temperature of 70 °C, and duration of 180 min.
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Figure 7. The effect of leaching duration on the recovery degrees of Fe, Cu, and Zn from the CBWS sample at the consumption rates of 13.26 mmol H2SO4/g and 2.94 mmol H2O2/g, temperature of 70 °C, and S/L ratio of 0.2 g/cm3.
Figure 7. The effect of leaching duration on the recovery degrees of Fe, Cu, and Zn from the CBWS sample at the consumption rates of 13.26 mmol H2SO4/g and 2.94 mmol H2O2/g, temperature of 70 °C, and S/L ratio of 0.2 g/cm3.
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Figure 8. The XRD pattern of the solid residue obtained under the optimal leaching conditions from the CBWS sample.
Figure 8. The XRD pattern of the solid residue obtained under the optimal leaching conditions from the CBWS sample.
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Metals 15 00330 g009aMetals 15 00330 g009b
Figure 9. SEM microphotograph of a random area of the solid residue obtained under the optimal leaching conditions of the CBWS sample (a) and its EDX mapping images (b–j).
Figure 9. SEM microphotograph of a random area of the solid residue obtained under the optimal leaching conditions of the CBWS sample (a) and its EDX mapping images (b–j).
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Figure 10. SEM microphotograph of some particles (a–c) of the solid residue obtained under the optimal leaching conditions from the CBWS sample and identified points for EDX analysis.
Figure 10. SEM microphotograph of some particles (a–c) of the solid residue obtained under the optimal leaching conditions from the CBWS sample and identified points for EDX analysis.
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Figure 11. Influence of oxidant type on iron, copper, and zinc recovery degrees under the following leaching conditions from the CBWS sample: acid consumption of 13.26 mmol H2SO4/g, S/L ratio of 0.2 g/cm3, temperature of 70 °C, and duration of 180 min.
Figure 11. Influence of oxidant type on iron, copper, and zinc recovery degrees under the following leaching conditions from the CBWS sample: acid consumption of 13.26 mmol H2SO4/g, S/L ratio of 0.2 g/cm3, temperature of 70 °C, and duration of 180 min.
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Figure 12. The effect of MnO2 consumption on the recovery degrees of Fe, Cu, and Zn from the CBWS sample at the consumption rate of 13.26 mmol H2SO4/g, S/L ratio of 0.2 g/cm3, temperature of 70 °C, and duration of 180 min.
Figure 12. The effect of MnO2 consumption on the recovery degrees of Fe, Cu, and Zn from the CBWS sample at the consumption rate of 13.26 mmol H2SO4/g, S/L ratio of 0.2 g/cm3, temperature of 70 °C, and duration of 180 min.
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Figure 13. Temperature dependence of free Gibbs energy change in Reactions (30)–(35).
Figure 13. Temperature dependence of free Gibbs energy change in Reactions (30)–(35).
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Table 1. The chemical composition of the CBWS sample, wt.%.
Table 1. The chemical composition of the CBWS sample, wt.%.
FeCuZnCCaSiAlCrMgMnNiCdAsPSPbSbBaTi
26.230.820.8117.18.995.402.800.404.402.000.0660.0280.340.162.200.300.060.200.14
Table 2. The compound composition of the CBWS sample, wt.%.
Table 2. The compound composition of the CBWS sample, wt.%.
α-FeFeOγ-Fe2O3FeOOHCCa2Al2SiO7Ca2MgSi2O7Mg2SiO4CaMgSiO4MgOCaSMnSFeAsCaTiO3
16.002.023.079.6217.110.279.793.004.102.832.183.170.590.40
ZnOZnSZn2SiO4ZnFe2O4CuNaAlSiO4Ca3MgSi2O8Pb2SiO4NiCrPbOBaOP2O5Sb2O3KAlSi2O6
0.550.200.190.460.823.342.000.330.0630.400.030.220.370.071.17
Table 3. Concentrations in diluted leaching solution and recovery degrees from the CBWS sample of main elements under the best leaching conditions.
Table 3. Concentrations in diluted leaching solution and recovery degrees from the CBWS sample of main elements under the best leaching conditions.
ElementsFeCuZnCaSiAlMgMnBaCdCrNiPbSbPTi
Concentration, mg/dm34072.41251171944.11945732900.0314.125.38.02.91.619.115.8
Recovery degree, %96.187.086.920.00.580.593.292.80.191.048.370.66.134.390.671.4
Table 4. Chemical composition of leaching residue obtained under the optimal conditions from the CBWS sample, wt.%.
Table 4. Chemical composition of leaching residue obtained under the optimal conditions from the CBWS sample, wt.%.
CaSSiCFeCuZnAlBaCdCrNiMgMnPPbSbTi
11.4912.76.0025.91.630.170.170.870.320.0040.330.0310.480.230.0240.450.0630.064
Table 5. Elemental composition analyzed by EDX of points 1–3 from the Figure 10, wt.%.
Table 5. Elemental composition analyzed by EDX of points 1–3 from the Figure 10, wt.%.
No.PhaseComposition, at. %
FeSZnCuCaSiO
1Iron sulfide30.669.4-----
2Iron–copper sulfide26.645.30.511.50.40.814.9
3Zinc sulfide2.252.642.3-0.92.0-
Table 6. Temperature dependencies of free Gibbs energy of Reactions (30)–(35).
Table 6. Temperature dependencies of free Gibbs energy of Reactions (30)–(35).
Reaction NumberReactionTemperature (K) Dependence of ΔG, kJ/mol
(30)Cu + 2Fe3+ = Cu2+ + 2Fe2+ΔG = −0.218∙T − 78.132
(31)Fe2+ + 0.5H2O2 + H+ = Fe3+ + H2OΔG = 0.173∙T − 100.131
(32)Fe2+ + 0.5MnO2 + 2H+ = Fe3+ + 0.5Mn2+ + H2OΔG = 0.165∙T − 47.358
(33)Cu + 0.5O2(g) + 2H+ = Cu2+ + H2OΔG = 0.160∙T − 176.077
(34)Fe + Cu2+ = Fe2+ + CuΔG = 0.004∙T − 156.698
(35)Fe + 2Fe3+ = 3Fe2+ΔG = −0.214∙T − 234.830
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Grudinsky, P.; Vasileva, E.; Dyubanov, V. Recycling Potential of Copper-Bearing Waelz Slag via Oxidative Sulfuric Acid Leaching. Metals 2025, 15, 330. https://doi.org/10.3390/met15030330

AMA Style

Grudinsky P, Vasileva E, Dyubanov V. Recycling Potential of Copper-Bearing Waelz Slag via Oxidative Sulfuric Acid Leaching. Metals. 2025; 15(3):330. https://doi.org/10.3390/met15030330

Chicago/Turabian Style

Grudinsky, Pavel, Ekaterina Vasileva, and Valery Dyubanov. 2025. "Recycling Potential of Copper-Bearing Waelz Slag via Oxidative Sulfuric Acid Leaching" Metals 15, no. 3: 330. https://doi.org/10.3390/met15030330

APA Style

Grudinsky, P., Vasileva, E., & Dyubanov, V. (2025). Recycling Potential of Copper-Bearing Waelz Slag via Oxidative Sulfuric Acid Leaching. Metals, 15(3), 330. https://doi.org/10.3390/met15030330

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