Zinc Kiln Slag Recycling Based on Hydrochloric Acid Oxidative Leaching and Subsequent Metal Recovery

Abstract

The limited availability of high-quality ore deposits and the environmental hazards of metallurgical wastes highlight the importance of developing resource-efficient metal recovery technologies. Zinc kiln slag (ZKS), also known as Waelz slag, a by-product material enriched in non-ferrous metals, was processed through oxidative HCl leaching with H₂O₂ as an oxidant. Thermodynamic simulation and laboratory experiments were applied to determine optimal leaching conditions to dissolve copper, zinc, and iron. Optimal leaching efficiency was achieved with consumptions of 0.8 g HCl and 0.1 g H₂O₂ per gram of ZKS, a liquid-to-solid (L/S) ratio of 5 mL/g, a temperature of 70 °C, and a duration of 180 min, which resulted in recoveries of 96.3% Cu, 93.6% Fe, and 76.8% Zn. The solid residue with 43.5 wt.% C is promising for reuse as a reductant material in pyrometallurgical processes. Copper and arsenic were separated from the leachate via cementation with iron powder, achieving recovery rates of 98.9% and 91.2%, respectively. A subsequent two-step iron precipitation produced ferric hydroxide with 52.2 wt.% Fe and low levels of impurities. As a result, the developed novel hydrochloric acid oxidative leaching and metal precipitation route for ZKS recycling provides an efficient and sustainable alternative to conventional treatment methods.

1. Introduction

The declining availability of high-grade raw materials is one of the key challenges for the modern non-ferrous metallurgical industry [1]. Such issues are noticeable in the zinc and copper metallurgical sectors, which suffer from a decrease in extractable primary ore reserves [2]. Additionally, the processing of these metals causes a significant environmental impact due to generating immense waste volumes [3] that are often improperly managed [4]. Zinc leach residue, copper smelter slag, zinc kiln slag, copper anode slime, alkaline and carbon zinc batteries are the main by-products with simultaneously elevated zinc and copper contents [5,6]. Depending on the source of the by-product, the residual amounts of zinc and copper in them vary over a wide range and can substantially exceed typical ore contents [7,8,9], making their extraction from these wastes reasonable. In particular, zinc kiln slag (ZKS), commonly known as Waelz slag, is an attractive material for the recovery of valuable metals due to their increased content [10].

ZKS is formed as a secondary product in the Waelz process, a widely applied roasting technology for the treatment of zinc-bearing materials using a rotary kiln [11]. About 60–80% of the treated feed is converted into kiln slag as a result of this process [12]. The feed can include zinc leach residue [13], electric arc furnace dust [14], dust from the copper industry [15], zinc oxide ores [16], and lead-bearing slag [10], which results in the chemical composition of ZKS being highly variable. ZKS usually contains such valuable elements as zinc and copper in the ranges of 0.6–2.1% and 0.2–3%, respectively, while iron is the primary element [17,18]. Moreover, significant amounts of silver and gold [19], as well as gallium and germanium [20], can be present, enhancing the value of ZKS as a raw material for the metallurgical industry.

The recycling of ZKS poses a challenge due to its complex composition that includes oxides, sulfides, and silicates of numerous metals [21]. The utilization of ZKS in diverse applications in construction [22,23] and ceramic [24] production have been proposed, but the high leachability of heavy non-ferrous metals often restricts such recycling approaches [25]. The recovery of valuable elements from ZKS provides substantial benefits, addressing both raw material limitations and environmental concerns. Several techniques have been used to extract valuable metals from ZKS [26]. Pyrometallurgical methods, including various roasting [27] and smelting [16,28,29] approaches, have been applied, but high energy consumption [30] makes them hardly efficient. Innovative processes such as autoclave ammoniacal leaching [31] and bioleaching [32] have been investigated for the recovery of valuable metals under environmentally mild conditions. However, high capital and operating costs [33] in the case of autoclave ammoniacal leaching, as well as a low leaching rate and limited effectiveness due to the presence of refractory minerals in ZKS [34], render these processes impractical. Other hydrometallurgical approaches, such as atmospheric leaching using sulfuric or hydrochloric acid, have been frequently utilized to recover valuable metals from various wastes [35,36,37] due to their efficiency for the selective dissolution of target metals. Hydrochloric acid leaching has been studied for the processing of magnetic iron concentrate derived from ZKS [38,39], showing itself as a promising method due to the high recovery rates of valuable elements.

Touching upon general hydrometallurgical challenges in valuable metal extraction, recent studies have increasingly focused on sustainable stepwise flowsheets for the recovery of Zn, Pb, In, and Ag from complex residues based on hydrochloric acid or chloride leaching routes [40,41], as well as on hybrid organic–chloride systems [42] to reduce reagent consumption and environmental impact. However, a systematic evaluation of such approaches for ZKS, whose complex mineral composition and high iron content complicate acid leaching efficiency, is still limited. This study addresses this gap by investigating hydrochloric acid oxidative leaching of copper-bearing ZKS with particular attention to the influence of various leaching parameters on the efficiency of non-ferrous metal recovery, as well as by proposing an integrated route for downstream recovery of Cu and Fe from the leachate through cementation and oxidation–neutralization. Therefore, to identify conditions that enable efficient and selective extraction of non-ferrous metals, the effects of different factors, such as acid concentration, liquid-to-solid (L/S) ratio, temperature, and leaching time, on the recovery rates of zinc, copper, and iron were systematically investigated. The leaching process yielded a solid residue, which was characterized to evaluate its recyclability. Copper and iron were recovered from the leachate using cementation with iron powder and oxidation–neutralization approaches, respectively, and the characteristics of the precipitated products containing Cu and Fe were assessed. The prospects of HCl leaching for the recovery of valuable elements from ZKS were evaluated.

2. Materials and Methods

2.1. Raw Materials

Table 1. The chemical composition of the ZKS sample, wt. %.

Fe	Ca	Si	Al	Mg	Mn	Na	K	P	S	Zn	Cu	Pb	As	Sb	Ni	Ti	V	Cr	Ba	C
26.23	8.99	5.40	2.80	4.40	2.00	0.54	0.21	0.16	2.20	0.81	0.82	0.30	0.34	0.06	0.066	0.14	0.01	0.40	0.20	17.1

and Figure 1 show the elemental and phase composition in the XRD pattern of the initial ZKS sample, respectively. The chemical composition confirms that iron is the major element in the ZKS sample. Minor elements such as Cu and Zn make the composition of the ZKS sample valuable for their extraction. The X-ray diffraction (XRD) pattern of the ZKS, identified using Match! 3.16 software (Crystal Impact, Bonn, Germany) [43] with the Crystallography Open Database (COD) database [44], reveals a complex mineralogical composition with iron-containing phases such as wuestite Fe_0.88O, metallic Fe, magnetite Fe₃O₄, and goethite FeOOH, as well as gangue minerals including melilite Al_0.5Mg_0.75Ca₂Si_1.75O₇, monticellite CaMgSiO₄, merwinite Ca₃MgSi₂O₈, and diopside Mg₂SiO₄. The detected graphite corresponds to unreacted coke fines from the pyrometallurgical process. The combination of complex iron phases and high carbon content reflects incomplete combustion of the reductants.

The following chemical reagents were used in the study: 37% H₂O₂ (Rushim, Moscow, Russia), >99.99% carbonyl Fe −0.05 mm powder (P/Ya M-5168, Torzhok, Russia), >97.2% FeCl₃ (Rushim, Moscow, Russia), and >97% Ca(OH)₂ (Antrazoxrom, Moscow, Russia).

2.2. Thermodynamic Simulation Method

Thermodynamic simulation was carried out using the Equilibrium Composition module of the HSC Chemistry 9.9 software (Metso, Helsinki, Finland) [45] at an atmospheric pressure of 100% nitrogen in the temperature range of 25–90 °C. We assumed an initial N₂ atmosphere instead of air to exclude the influence of atmospheric oxygen on the quantitative evaluation of the equilibrium composition for revealing the specific effect of H₂O₂ as an oxidizing agent. The initial amount of ZKS was 100 kg. The initial mass ratio of liquid to solid was always equal to 5. In the simulation, the gas phase was assumed to be as an ideal gas. All solid components were considered pure phases. The aqueous phase was calculated as an ideal solution for the majority of species or using the Pitzer model [46] for copper and zinc species, which achieves the closest correspondence between the modeling and experimental data. The study [47] confirmed difficulties in the simulation of the aqueous phase using the Pitzer model for ions with high concentrations, where the indicated maximum level of ion concentration for an accurate calculation was 6 mol/kg.

Taking into account the results of the chemical and phase analyses of the ZKS sample (Section 2.1), its model composition was adopted for thermodynamic calculations as follows (%): 10.27 Ca₂MgSi₂O₇; 9.79 Ca₂Al₂SiO₇; 16.7 Fe; 3.2 γ-Fe₂O₃; 2.11 FeO; 10.08 FeOOH; 17.1 C; 3.17 MnS; 0.46 ZnFe₂O₄; 4.10 CaMgSiO₄; 2 Ca₃MgSi₂O₈; 3 Mg₂SiO₄; 2.83 MgO; 0.55 ZnO; 0.2 ZnS; 0.19 Zn₂SiO₄; 0.74 Cu; 2.18 CaS; 3.34 NaAlSiO₄; 1.17 KAlSi₂O₆; 0.4 Cr; 0.33 Pb₂SiO₄; 0.03 PbO; 0.37 P₂O₅; 0.59 FeAs; 0.07 Sb₂O₃; 0.22 BaO; 0.4 CaTiO₃; 0.07 Ni.

2.3. Experiments

2.3.1. Hydrochloric Acid Leaching of the ZKS Sample

The leaching trials were carried out utilizing a heated magnetic stirrer model HJ-4B (Changzhou Surui Instrument Co., Ltd., Changzhou, China) with a rotating bar to maintain agitation. The experimental setup involved introducing 2–8 g of the ZKS sample into a 50 mL conical flask. A precisely measured volume, ranging from 12.8 to 32 mL, of HCl at the required concentration was then added. The suspension was continuously mixed using a Teflon-coated stirring bar. Under optimized conditions, the scale of the experiment was enlarged tenfold.

The temperature regulation was managed via a thermocouple housed in a quartz sheath, which was immersed in the reaction mixture. The investigation focused on analyzing the impact of various parameters, including HCl consumption (0.7–1 g per gram of ZKS), H₂O₂ consumption (0–0.25 g per gram of ZKS), operational temperature (25–90 °C), initial liquid-to-solid ratio (2–20), and holding time (5–180 min) on the efficiency of zinc, copper, and iron recovery. The stirring rate remained constant at 400 rpm. Each leaching test was conducted once per experimental condition.

If necessary, the solution was blown by air during the stirring using a silicone hose with several holes 2 mm in diameter inserted through the neck of the flask. An air tank was used to supply the gas into the flask. In order to control a constant gas flow rate, rotameters were integrated into the circuit between the tank and the flask. The gas flow rate was 0.5 L/min.

After processing under the specified conditions, the solid phase and leachate were separated via vacuum filtration. The residual solid was rinsed with acidified distilled water adjusted to pH = 2, followed by drying at 90 °C for a retention time of 120 min. The obtained filtrate was subsequently diluted to a final volume of 200 or 250 mL using acidified (pH = 1) distilled water.

The recovery rate of elements after hydrochloric acid leaching was computed using the following equations:

$$\%\omega ={\displaystyle \frac{V_M·X_L}{m_0·{\%X}_0}}·100$$

(1)

$$\%\omega =100-{\displaystyle \frac{m_{SR}·{\%X}_{SR}}{m_0·{\%X}_0}}·100$$

(2)

where %ω—recovery rate of elements into the leachate, %; m₀—weight of the ZKS sample, g; %X₀—content of an element in the ZKS sample, %; X_L—concentration of an element in the leachate, g/L; V_M—volume of a measured flask, L; %X_SR—concentration of an element in the residue, %; m_SR—mass of the residue, g.

The percentage of the element remaining in the residue or precipitates was determined accordingly:

$$\%\beta ={\displaystyle \frac{m_{SR}·{\%X}_{SR}}{m_0·{\%X}_0}}·100$$

(3)

where %β—percentage of elements remaining in the residue, %.

2.3.2. Precipitation of Copper and Iron

In order to recover copper, the leachate was treated using cementation with high-purity iron powder under heated stirring conditions. A pre-weighed amount of 0.272 g of iron powder was introduced into a 100 mL conical flask, and the suspension was agitated at 40 °C and a stirring speed of 400 rpm for 1 h. A dark, non-magnetic solid phase gradually formed during the process. After the processing, any residual magnetic material adhering to the stirrer was removed, and the resulting precipitate containing copper was separated from the liquid phase by vacuum filtration. The collected filtrate was diluted to 100 mL and subjected to an iron precipitation stage.

In one step, iron removal from the solution was carried out by first promoting the oxidation of ferrous ions to the ferric state through the controlled addition of 12.4 M H₂O₂ at elevated temperature. A twofold excess of the oxidizing agent was introduced dropwise at approximately 70 °C. After the oxidation, the pH of the system was gradually increased using a freshly prepared suspension of Ca(OH)₂ (10 wt.% in water). The pH level was continuously monitored using an immersed pH-meter. Upon reaching a target pH = 3.6, a ferric hydroxide precipitate formed, which was separated by vacuum filtration and subsequently washed twice with slightly acidified water adjusted to pH = 3.

A two-stage approach was applied for iron removal, involving partial oxidation of ferrous species followed by sequential precipitation. Initially, approximately 30% of the theoretical amount of H₂O₂ was introduced to partially convert Fe²⁺ to Fe³⁺. This was followed by pH adjustment to 3.6 using the Ca(OH)₂ suspension. The resulting ferric precipitate was separated from the liquid phase by vacuum filtration without further washing. The filtrate was then acidified with HCl to pH = 2 and reheated up to 70 °C. At this stage, a twofold excess of H₂O₂ was added to complete the oxidation of the remaining ferrous ions. Once oxidation was complete, the pH was again raised to 3.6, leading to the formation of additional ferric hydroxide. The solid phase from the second step was also collected by vacuum filtration and washed twice with slightly acidified water (pH = 2).

All the obtained precipitates were dried at 90 °C for 2 h and subsequently weighed for mass balance calculations.

Figure 2 visualizes the described methods of copper and iron precipitation in a simplified diagram.

2.4. Analysis Methods

Inductively Coupled Plasma Optical Emission Spectroscopy (ICP-OES) was applied to measure the concentrations of iron, zinc, copper, and other elements in the leachates using a Varian Vista Pro (Varian Optical Spectroscopy Instr., Mulgrave, Australia) spectrometer. Moreover, the ZKS sample was also analyzed by ICP-OES after fusion with a mixture of sodium carbonate and borax at 950 °C with a subsequent digestion in 1 M HCl, whereas the preparation of iron precipitates for ICP-OES analysis involved only the digestion with concentrated HCl.

Carbon and sulfur concentrations in the ZKS and leach residue samples were analyzed using a LECO CS-230 analyzer (LECO Corporation, St. Joseph, MI, USA). Other elements in the leach residue were analyzed by X-ray fluorescence spectrometer AXIOSmax Advanced (PANalytical, Almelo, The Netherlands).

Scanning electron microscopy (SEM) was employed to analyze the obtained residues and precipitates using a Tescan Vega 3SB microscope (Tescan, Brno, Czech Republic) integrated with an INCA SDD X-MAX (Oxford Instruments, Abingdon, UK) detector, which uses energy dispersive X-ray spectroscopy (EDS) for local chemical analysis.

XRD patterns were recorded on a Tongda TDM-20 diffractometer (Dandong Tongda Science & Technology Co., Ltd., Dandong, China) using Cu K_α radiation. Phase analysis was performed with Match! 3.16 software supported by the COD.

3. Results

3.1. Thermodynamic Simulation Results

Figure 3 shows the effect of HCl addition on equilibrium amounts of Fe, Mn, Ca, Mg, Si, Zn, and Cu compounds at 25 °C, where the added amount of HCl and its corresponding concentration are indicated on the bottom and top horizontal axes, respectively.

As the amount and concentration of HCl rise in Figure 3a, solid phases disappear in a clear sequence: calcium and magnesium silicate hydrates (3CaO·Al₂O₃·3SiO₂, 3MgO·2SiO₂·2H₂O, and 3MgO·4SiO₂·H₂O) are consumed, releasing Ca²⁺ and Mg²⁺ while leaving a SiO₂ residue. Next, MnS dissolves, producing Mn²⁺. Finally, Fe₃O₄ and FeOOH are dissolved, giving a rise in Fe²⁺ ions. It was calculated that at least 77 kg of HCl for 100 kg of ZKS (0.77 g/g) is needed for full dissolution of Fe compounds into Fe²⁺. After that, the solution is dominated by Fe²⁺, Ca²⁺, Mg²⁺, Mn²⁺ and SiO₂. The presence of a minor amount of elemental sulfur also should be noted. According to the calculations in Figure 3b, ZnS remains almost completely undissolved across the whole acid range, and Zn²⁺ appears only once the surplus of HCl is achieved. Metallic Cu is also stable until HCl amounts exceed about 90 kg; at that point it is converted into CuCl, which is also poorly soluble as well. Thus, Figure 3b demonstrates that the dissolution of Cu with HCl demands additional reagents.

In order to dissolve copper, air blowing as well as the addition of H₂O₂ or an Fe³⁺-containing solution during leaching were considered. These oxidizing agents are well known in copper hydrometallurgy [48,49]. Since thermodynamic calculations cannot provide essential information regarding the reaction kinetics, preliminary experiments under identical conditions were conducted to evaluate the performance of the oxidants. Figure 4 shows the results of preliminary experiments comparing the effect of the selected oxidizing agents on the recovery rate of Fe, Cu, and Zn. As follows from the presented data, the recovery rate of Fe remains high and differs insignificantly under all experimental conditions. The recovery rate of Zn is in the range of 76–84%, which also indicates an insignificant difference in the values. In the absence of oxidizing agents, copper does not pass into the leachate under hydrochloric acid leaching conditions, as predicted by the thermodynamic calculations (see Figure 3). Air blowing increases the recovery rate of Cu up to about 65%, while the addition of Fe³⁺ ions raises it only up to about 21%. In contrast, the addition of H₂O₂ increases copper recovery rate up to 92%. Thus, H₂O₂ was used for the following calculations and experiments as the most effective oxidizer for the copper recovery.

Figure 5 shows the thermodynamic modeling results for the equilibrium amounts of Fe, Mn, Ca, Mg, Si, Zn, and Cu compounds influenced by H₂O₂ amounts at 25 °C with an excess addition of HCl to 100 kg of ZKS.

Figure 5a demonstrates that as the H₂O₂ amount increases, Fe²⁺ is gradually oxidized to Fe³⁺ according to reaction (4) with a complete conversion occurring around 25 kg H₂O₂:

H₂O₂ + 2Fe²⁺ + 2H⁺ → 2H₂O + 2Fe³⁺

(4)

This process is crucial for Cu dissolution according to reaction (5), which is displayed in Figure 5b:

Cu + Fe³⁺ → Cu²⁺ + Fe²⁺

(5)

In addition, Figure 5a shows a slight decrease in Ca²⁺ concentration after about 10 kg H₂O₂ due to the precipitation of gypsum (CaSO₄·2H₂O). SiO₂ remains inert despite H₂O₂ addition. The Mn²⁺ and Mg²⁺ concentrations are ultimately unchanged, suggesting they are unaffected by H₂O₂ addition under these conditions.

In Figure 5b, Cu and ZnS are stable at low H₂O₂ amounts, but both dissolve sharply at 17 kg of H₂O₂, leading to the formation of Cu²⁺ and Zn²⁺ ions in solution. The addition of H₂O₂ as oxidizer was proven to be crucial for Cu dissolution.

Figure 6 exhibits further investigation of the optimal conditions: it displays thermodynamic modeling of the temperature effect on Fe, Mn, Ca, Mg, Si, Zn, and Cu compounds with the addition of an excess of HCl and H₂O₂ to 100 kg of ZKS.

According to the calculations shown in Figure 6a, CaSO₄·2H₂O fully decomposes into CaSO₄ at around 40 °C. The calculations also show profuse evaporation of HCl at higher temperatures, which can lead to the precipitation of Fe₂O₃. Mn²⁺, Mg²⁺, SiO₂, and Ca²⁺ remain relatively stable across the temperature range.

In Figure 6b, Zn²⁺ remains nearly constant, indicating that temperature has little effect on its solubility. Cu²⁺ is stable up to 80 °C, after which it sharply decreases due to the formation of Cu⁺ species, suggesting partial reduction.

In summary, according to the thermodynamic simulation, the required HCl and H₂O₂ consumptions were estimated to be 0.77 and 0.25 g/g ZKS, respectively. In order to make a round value of the HCl consumption, we assume 0.8 g/g for verification in subsequent experiments.

3.2. Leaching Experiments

Based on the thermodynamic calculations, experimental tests were carried out under conditions close to the calculated HCl and H₂O₂ consumption values and temperature to verify the theoretical predictions and determine the optimal leaching parameters.

In order to optimize the oxidant addition, we firstly tested different H₂O₂ consumption rates. Figure 7 illustrates the influence of H₂O₂ consumption rate on the recovery rates of Fe, Cu, and Zn under fixed conditions.

A rapid change in the Cu recovery rate was noticed with the addition of 0.05 g/g H₂O₂, while its highest value of 92.1% was reached at 0.1 g H₂O₂/g. Zinc and iron also reach their maximum recoveries at 0.1 g H₂O₂/g with their recovery rates of 83.6 and 96.6%, respectively. In contrast to the thermodynamic calculations (Figure 5), a further increase in H₂O₂ consumption rate up to 0.25 g/g led to a gradual decrease in the element recovery rates, so 0.1 g H₂O₂/g was chosen as the optimal oxidant consumption.

Figure 8 illustrates the experimental results of leaching using various HCl consumptions both without and with hydrogen peroxide addition.

Figure 8a focuses on the recovery rates of Fe, Cu, and Zn depending on HCl consumption rate with the corresponding HCl concentration at indicated on the top horizontal axis at the chosen conditions based on the thermodynamic simulation (Figure 3a,b). The Fe curve indicates the highest recovery out of all elements reaching its maximum value at 1 g HCl/g. Zn recovery levels fluctuate between 70 and 80%. Cu recovery rates remain below 2%, which unambiguously proves that Cu remains undissolved using only HCl. Figure 8b demonstrates the influence of HCl consumption rate on the recovery rates of Fe, Cu, and Zn with the addition of 0.1 g H₂O₂/g. Cu recovery rate shows a rapid increase with the addition of 0.8 g/g HCl, proving the crucial role of H₂O₂ for Cu leaching efficiency. Fe recovery rate continues to grow with the increase in HCl consumption rate, although Zn recovery rates at 0.8 and 0.9 g HCl/g consumption rates stay roughly the same. Thus, 0.8 g HCl/g was considered the optimal consumption rate for further experiments.

Figure 9 depicts the effect of temperature, L/S ratio, and leaching time on recovery rates of copper, zinc, and iron with the addition of 0.8 g HCl/g and 0.1 g H₂O₂/g.

Figure 9a presents the effect of temperature on the recovery rates of Fe, Cu, and Zn. Whereas Fe recovery rates remain consistently high, the recovery rates of Cu and Zn slightly increase under higher temperatures. With Cu recovery being top priority and it peaking at 70 °C, this temperature was chosen to be the optimum for further experiments, in spite of Zn recovery rate reaching its highest level at 60 °C. A decrease in Fe and Cu recovery rates at 80–90 °C may be attributed to HCl evaporation (Figure 6a) at elevated temperatures, which lowers the acid concentration and, consequently, reduces the solubility of the metals [50].

Figure 9b displays the recovery rates of the elements as a function of the L/S ratio. The copper recovery rate sharply grows within the L/S ratio range of 2.25–5 mL/g and achieves the highest value at 5 mL/g. The Fe and Zn recovery rates also reach their maximum values at this level, so the optimal L/S ratio was chosen to be 5 mL/g.

Figure 9c demonstrates the impact of leaching time on the recovery rates of Fe, Cu, and Zn. The recovery rates of Fe, Cu, and Zn show slight fluctuations at smaller leaching periods; however, as it is obvious from the plot, the highest recovery rates of all three elements are achieved after 180 min of leaching.

In summary, the optimal leaching conditions were established as 0.8 g HCl/g and 0.1 g H₂O₂/g consumption rates, an L/S ratio of 5 mL/g, a temperature of 70 °C, and a leaching time of 180 min, under which 96.3% Cu, 93.6% Fe, and 76.8% Zn were recovered.

3.3. Characterization of Leaching Products

Table 2. Chemical composition of the HCl leach residue, as well as the element recovery rates into the residue and into the leachate, respectively, wt. %.

Element	Chemical Composition, wt. %	Remaining in the Residue (%β)	Recovery into the Leachate (%ω)
Fe	5.89	8.6 ± 2.3	91.4 ± 2.3
Cu	0.35	16.3 ± 4.5	83.7 ± 4.5
Zn	0.34	15.9 ± 4.5	84.1 ± 4.5
Al	1.28	17.4 ± 2.6	82.6 ± 2.6
As	0.49	55.3 ± 2.4	44.7 ± 2.4
Ba	0.37	70 ± 1.6	30 ± 1.6
Ca	1.00	4.3 ± 3.1	95.7 ± 3.1
Cd	0.0039	7.1 ± 3.0	92.9 ± 3.0
Cr	0.73	69.5 ± 1.6	30.5 ± 1.6
K	0.43	77.3 ± 1.2	22.7 ± 1.2
Mg	0.64	5.6 ± 5.1	94.4 ± 5.1
Mn	0.36	6.9 ± 5.0	93.1 ± 5.0
Na	0.16	11.3 ± 4.8	88.7 ± 4.8
Ni	0.008	4.6 ± 9.7	95.4 ± 9.7
P	0.355	84.5 ± 0.8	15.5 ± 0.8
Pb	0.064	8.1 ± 9.4	91.9 ± 9.4
Sb	0.11	71.7 ± 1.5	28.3 ± 1.5
Si	12.07	85.1 ± 0.5	14.9 ± 0.5
Ti	0.36	98.8 ± 0.07	1.2 ± 0.07
C	43.5	96.9 ± 0.1	3.1 ± 0.1
S	4.0	69.3 ± 0.98	30.7 ± 0.98
Cl	2.19	n/a 1	n/a

¹ n/a—not applicable.

lists the chemical composition of the leach residue obtained from 40 g of ZKS and the element recovery rates under optimal leaching conditions. The yield of the residue was 15.2385 g or 38.1%.

The chemical composition of the residue indicates that Fe, Cu, Zn, Al, Cd, Mg, Mn, Na, Ni, Pb, and Ca were mainly dissolved with recovery rates of 80–96% into the leachate due to the formation of highly soluble chlorides, which is consistent with the thermodynamic predictions (Figure 3). It is noteworthy that scaling up the experiment changed marginally the recovery rates of Cu, Fe, and Zn. Specifically, the recovery rates of Cu and Fe decreased from 96.3 to 83.7% and from 93.6 to 91.4%, respectively, whereas the recovery rate of Zn increased from 76.8 to 84.1%. This effect can be attributed to scale-dependent changes in leaching kinetics and mass-transfer conditions, which influence the dissolution behavior of different elements [51,52]. In contrast, As, Sb, S, Ba, K, and Cr exhibited only partial dissolution, while Si, Ti, P, and C were almost completely retained in the solid phase.

In order to elucidate the reasons of a behavior of various elements and investigate the leach residue in more detail, the XRD pattern and SEM images were thoroughly analyzed. Figure 10 demonstrates the XRD pattern of the residue with identified phases.

XRD analysis revealed the presence of graphite C, such spinel phases as Al₂MgO₄ and Cr₂FeO₄, orthorhombic elemental sulfur S₈, gypsum CaSO₄·2H₂O, orpiment As₂S₃, and complex sodium-potassium chloride Na_0.6K_0.4Cl. Hence, the insolubility of carbon is mainly associated with its presence as chemically inert graphite. The partial retention of Al, Mg, Cr, and Fe in the residue is attributed to their presence in the form of insoluble spinel phases MgAl₂O₄ and FeCr₂O₄, which are resistant to hydrochloric acid leaching [53,54,55]. The formation of S₈ during leaching, as was predicted by thermodynamic simulation (Figure 3a), is caused by the release of H₂S from the reaction of sulfides with HCl, followed by its oxidation by dissolved oxygen to elemental sulfur [56,57]. The presence of gypsum and complex sodium-potassium chloride in the residue apparently results from secondary crystallization from the sulfate- and chloride-bearing mother liquor during washing and drying. The formation of orpiment As₂S₃ in hydrochloric acid solutions occurs when acid-soluble arsenic reacts with H₂S released from sulfide dissolution [58]. In addition, the diffraction profile and Table 2 indicate that SiO₂ is present in a non-crystalline state, being the second most abundant phase after graphite.

Figure 11 clearly shows SEM images of the leach residue, whereas

Table 3. Chemical composition of the marked points in Figure 11 using EDS analysis, at. %.

Element	Spectrum Point
	1	2	3	4	5	6
C	98.8	-	-	-	-	-
Si	0.6	28.8	3.9	-	1.4	-
S	0.3	2.2	87.3	69.2	48.6	67.0
Ti	0.2	0.2	-	-	-	-
Al	0.1	0.7	-	-	-	-
O	-	60.5	8.1	-	3.1	-
Fe	-	3.4	0.3	30.8	5.1	-
Cl	-	2.2	-	-	-	-
Cr	-	0.2	-	-	-	-
Ba	-	0.2	-	-	-	-
Cu	-	0.2	-	-	-	33.0
As	-	0.2	-	-	-	-
Mg	-	0.5	-	-	-	-
P	-	0.5	-	-	-	-
Ca	-	-	0.3	-	-	-
Zn	-	-	-	-	40.1	-
Mn	-	-	-	-	1.6	-

indicates the composition of marked points in the images.

As follows from Figure 11a and Table 3, local SEM-EDS microanalysis confirmed the presence of the major phases: graphite, non-crystalline silica, and elemental sulfur. Silica envelops other particles, which results in an elevated silicon content in other areas of the sample.

According to Figure 11b–d, residual iron, zinc, and copper were detected in the form of various sulfides. Considering this, the gradual dissolution of Zn (Figure 9c) is likely attributed to the presence of a sulfide fraction, which dissolves slowly in HCl as a result of the development of a passivation layer during leaching [59]. Iron was identified as pyrite (Table 3), which is poorly soluble in HCl and can only be partially dissolved under moderately oxidative hydrochloric acid leaching conditions [60]. Table 3 also indicates the presence of copper in a pyrite-like disulfide form, CuS₂, which is likely the ultimate oxidation product of Cu₂S during leaching, while elemental sulfur is generated at the final stage of this process [61].

In summary, the HCl leach residue appears to be graphite-based material with a significant proportion of non-crystalline SiO₂ and various minor metal compounds.

3.4. Copper Precipitation

Table 4. Chemical composition of the HCl leachate before and after cementation, and the recovery rates of key elements, wt. %.

Element	Composition of Initial Mother Leachate, mg/L (V = 50 mL)	Composition of Solution After Cementation, mg/L (Dilution to 100 mL)	Recovery from the Leachate into the Cementation Product, %
Fe	28,090	15,200	−6.8 1
Cu	830	4.5	98.9
Zn	788	330	16.2
As	102	4.5	91.2
Cd	23.1	11.5	0.43

¹ negative value (−6.8%) indicates partial dissolution of Fe powder during the cementation stage.

lists the chemical composition of the mother liquor before and after cementation with iron powder, as well as the cementation treatment efficiency for the recovery of copper, zinc, arsenic, and cadmium.

As shown from the listed data, cementation with iron powder resulted in different behaviors of the dissolved elements. Cu was almost completely removed from the leachate with the recovery rate of 98.9% that demonstrates a high efficiency of the cementation stage. Fe showed a negative recovery due to partial dissolution of the added iron powder and subsequent enrichment of the leachate. Zn and Cd showed the recovery rates of 16.2% and 0.43%, respectively, which indicates their predominant remaining in the dissolved form. Minor zinc removal occurred due to intensive Fe corrosion during the treatment, which produced fine ferric (oxy)hydroxides and flocculent phases that adsorb or co-precipitate Zn²⁺ [62]. An important result was the 91.2% of As removal due to its affinity with co-precipitation or sorption during the cementation process [63], which enabled the recovery of copper and the most of arsenic simultaneously.

Figure 12 exhibits a SEM image of a cementation residue particle and the distribution of elements within it.

Table 5. Chemical composition of the marked point in Figure 12 using EDS analysis, at. %.

Spectrum Point	Element
	Cu	Fe	Cl	O	S	Si
7	18.7	15.3	15.9	49.1	0.6	0.4

provides the chemical composition of the marked points in Figure 12.

As observed from the presented data, the non-magnetic copper-containing residue obtained after cementation points out an irregular, porous morphology with a rough surface and consists mainly of Cu, Fe, O, and Cl, suggesting the formation of mixed copper-iron hydroxides or oxides with chloride phases. The presence of local areas rich in Si and Ca in the residue (Figure 12f,g) is attributed to their adsorption onto ferric hydroxides and simultaneous precipitation [64,65].

Overall, the cementation stage effectively removed most of the Cu and As into the solid phase, and the obtained solution was subsequently used to study the iron precipitation process.

3.5. Iron Precipitation

Table 6. Chemical composition (CC) of ferric hydroxide precipitates derived from the one-step and two-step approaches, along with element recovery rates (%ω ± Δ%ω) from the initial mother leachate.

Approach	Fraction (Mass)	Index	Element
			Fe	Al	Cr	Sb	Mn	Zn	Cu	Pb	As	Cd	Ni
One-step	Single (100%)	CC, wt.%	36.1	1.52	0.11	0.002	0.42	0.63	0.02	0.30	0.043	0.006	0.027
		%ω	93.8	85.1	81.3	23.2	14.7	57.8	1.6	62.6	30.5	18.1	50.9
		±Δ%ω	2.7	2.4	2.3	0.7	0.4	1.6	0.04	1.8	0.9	0.5	1.4
Two-step	First (28.9%)	CC, wt.%	38.0	3.53	0.65	0.015	0.14	0.11	0.02	0.26	0.188	0.003	0.005
		%ω	20.8	41.6	97.1	40.7	1.0	2.1	0.3	11.2	28.3	1.9	1.8
		±Δ%ω	0.6	1.2	2.8	1.2	0.03	0.06	0.01	0.3	0.8	0.05	0.05
	Second (71.1%)	CC, wt.%	52.2	0.79	0.01	0	0.03	0.02	0.01	0.06	0.001	<0.001	0.001
		%ω	70.1	23.1	2.4	0	0.5	0.9	0.5	6.5	0.3	0.6	1.2
		±Δ%ω	2	0.7	0.07	0	0.014	0.026	0.015	0.185	0.008	0.017	0.033

compares the one-step and two-step approaches for the precipitation of ferric hydroxide. The presented data highlight a variety in the chemical composition of the obtained precipitates and the corresponding recovery rates of elements from the solution.

In the one-step approach, most of the Fe and major impurities such as Al and Cr were co-precipitated in a single fraction with recovery rates more than 85%. Al and Cr co-precipitate with ferric hydroxide because of forming insoluble hydroxides at similar pH values, as well as their incorporation through surface adsorption and co-precipitation within the amorphous ferric hydroxide matrix [66,67]. Moreover, just over half of Zn, Pb, and Ni was also precipitated with ferric hydroxide due to adsorption onto hydrous ferric oxides and carrier co-precipitation [68]. Other elements, such as Mn, Sb, Cu, and Cd, were co-precipitated inconsiderably. The remaining portion of As was precipitated from the solution after copper cementation due to a high efficiency of As adsorption and co-precipitation on freshly formed ferric hydroxide phases [69]. A discrepancy of the As element balance (Table 4 and Table 6) is likely attributed to arsenic introduced from the reagents used. In general, although the one-step approach shows the Fe recovery rate of 93.8%, it generates ferric hydroxide, which is difficult to recycle due to high contents of various impurities.

In contrast, the two-step procedure separates the process into two fractions with distinct chemical compositions and recovery rates. The first fraction captured 20.8% of Fe and the majority of impurities, whereas the second fraction recovered 70.1% of Fe. It contained impurities at significantly lower levels, in some cases, by up to two orders of magnitude. The difference in content is greater for elements that are strongly adsorbed onto ferric hydroxide, such as Cr and As. The second fraction of ferric hydroxide contains a high Fe percentage and negligible amounts of impurities, with the exception of Al, which accounts for 0.79 wt.%. As a result, the two-step approach indicates effective separation of impurities compared with the single-step method.

Figure 13 displays SEM images of both ferric hydroxide fractions obtained in the two-step approach, along with the compositions of the marked points listed in

Table 7. Chemical composition of the marked points in Figure 13 using EDS analysis, at. %.

Spectrum Point	Element
	Fe	Al	Ca	Mg	Na	Zn	Pb	As	P	Cr	Si	S	Cl	O
8	23.1	6.1	2.1	0.4	-	0.3	1.5	-	-	0.3	0.3	1.6	12.9	51.4
9	11.2	4.5	0.3	-	-	-	-	0.3	0.2	0.3	-	0.8	6.1	76.3
10	25.0	1.4	0.4	-	-	-	-	-	-	-	-	0.8	9.3	63.1
11	18.9	1.2	0.2	-	1.6	-	-	-	-	-	-	0.6	7.8	69.7

.

Figure 13a clearly shows heavy metal impurities within one of the particles of the first ferric hydroxide fraction as bright inclusions. The local analysis data (Table 7) are generally consistent with the bulk compositions of the hydroxides (Table 6). In addition, a significant amount of Cl and a substantial amount of S were detected. Chlorine from the HCl solution was probably incorporated into ferric hydroxide precipitates through the formation of chloride-bearing phases [70], while sulfur was likely adsorbed as sulfate ions onto ferric hydroxides and co-precipitated in basic ferric sulfate phases [71].

In summary, the two-step approach enables recovery of more than 90% of Fe in total as ferric hydroxide fractions, where about 70% of the recovered Fe is present as the fraction characterized by low impurity levels.

4. Discussion

Hydrochloric acid oxidative leaching demonstrated high efficiency in the extraction of valuable elements from ZKS, achieving recoveries of 96.3% Cu, 93.6% Fe, and 76.8% Zn. As shown in our previous work [72], these values are comparable to those obtained by sulfuric acid leaching of ZKS using hydrogen peroxide as an oxidant, achieving recoveries of 87.0% Cu, 96.1% Fe, and 86.9% Zn.

It should be clarified that the overall oxidation mechanism of copper in the HCl medium with hydrogen peroxide is well established: Fe³⁺ is generated from Fe²⁺ by H₂O₂ according to reaction (4), although the exact reaction pathway is still under discussion [73], while metallic Cu is dissolved through reaction (5) involving Fe³⁺ as the oxidant. As shown in Figure 4, compared with direct Fe³⁺ addition and air blowing, H₂O₂ proved to be the most efficient oxidizing agent owing to its higher oxidative potential and continuous regeneration of Fe³⁺ during leaching [49], which also results in faster kinetics (Figure 9c). The dissolution of copper can also be achieved using other oxidants such as sodium chlorate NaClO₃ [74] or ozone O₃ [75], which similarly oxidize Fe²⁺ to Fe³⁺ in acidic media. In addition, a reagent-saving alternative is electrochemical oxidation using a PbO₂ anode for the oxidation of Fe²⁺ to Fe³⁺ [76], avoiding the need for external oxidants. However, in all cases, it should be emphasized that copper dissolution in chloride media can be enhanced by Cu–Cl complex formation [77], which significantly accelerates leaching kinetics [78].

Although H₂SO₄ is traditionally employed in copper and zinc metallurgy, HCl also provides a distinction associated with its ability to dissolve calcium. As a result of sulfuric acid leaching, a calcium sulfate-based residue is typically formed, which can potentially be utilized in the construction industry [72]. In contrast, hydrochloric acid leaching generates a residue mainly composed of coke breeze carbon and amorphous silica (Figure 10). A potential drawback of the HCl leaching is the presence of a substantial As content (Table 2) in the residue, which may either accumulate in the system or volatilize during a subsequent recycling process, thereby causing environmental concerns. Partial precipitation of arsenic during leaching limits the recycling potential of the residue. Furthermore, the presence of sulfides in ZKS (Figure 11) complicates the dissolution of metals, which is consistent with the study [39], indicating that the addition of sulfide-rich materials to the Waelz kiln feed is undesirable if the subsequent leaching is applied. Taking into account this fact, as well as the complex and refractory phase composition of the ZKS sample (Figure 1), alkaline lixiviants and weak acids seem to be inevitably ineffective for the recovery of valuable elements. Only strong mineral acids can provide sufficient reactivity. Along with the widely used sulfuric and hydrochloric acids, the use of nitric acid and perchloric acid [79,80] has been reported for related materials. However, their industrial application is impractical because nitric acid generates nitrate-rich environmentally hazardous effluents [81], while perchloric acid poses severe safety, corrosion, and explosion risks. In contrast, hydrochloric acid offers an optimal balance of leaching efficiency and operational safety owing to its ability to effectively decompose refractory phases and produce chloride solutions that can be recycled within a closed hydrometallurgical circuit [82].

Nevertheless, this type of HCl residue offers alternative utilization routes, particularly as a reducing agent in pyrometallurgical operations. A primary option is its reuse in the Waelz process, where ZKS is produced. Undoubtedly, further treatment is desirable, namely, water washing to remove residual chlorides (Table 2, Figure 10). In addition, the carbon content can be substantially increased by froth flotation [83]. The addition of the leach residue to secondary zinc-bearing feed materials may reduce the consumption of coke breeze in the Waelz process, which constitutes one of the largest cost factors [84]. According to the waste management hierarchy [85], such in-process reuse is preferable to recycling in unrelated technological processes.

The recovery rate of 98.9% from the leachate achieved in the cementation stage is consistent with the results of other studies [86,87]. Reportedly, the process is characterized by low cost, ease of control, operational simplicity, and high efficiency [88]. It is important to note that cementation not only precipitates Cu together with most of As, but also neutralizes the leachate, thereby preparing it for the subsequent stage of Fe recovery. In order to further reduce processing costs, secondary iron materials such as cast iron turnings [89], steel scrap [90], and iron shavings [91] can be applied during cementation.

An important result is the removal of the main part of arsenic from the solution along with copper as copper arsenide Cu₃As [92]. The mechanism of As precipitation during cementation in the HCl medium was previously elucidated [93], proceeding according to the following reactions:

AsO⁺ + 2H⁺ + 3e → As + H₂O

(6)

As + 6Cu → Cu₃As + 3Cu⁺ + 3e

(7)

AsO⁺ + 2H⁺ + 6Cu → Cu₃As + H₂O + 3Cu⁺

(8)

Reactions (6) and (7) represent the elementary steps of the overall process described by reaction (8). In this mechanism, AsO⁺ ions in the leachate are first reduced to elemental arsenic according to reaction (6). The deposited arsenic subsequently reacts with metallic copper to form Cu₃As, as described by reaction (7).

The residual part of arsenic in the leachate after the cementation stage was effectively adsorbed onto ferric hydroxide (Table 5). The recovery of arsenic into the solid products is environmentally relevant, since it allows almost complete fixation of such a harmful contaminant as arsenic. Locking arsenic into solid, hardly soluble materials is more environmentally friendly than releasing it with liquid streams. This approach supports sustainable waste management because it keeps the toxic element safely contained, while allowing valuable metals to be recovered.

In contrast to the one-step approach for iron precipitation, the two-step approach appeared to be preferable due to the production of a ferric hydroxide product with low contents of impurities (Table 6). Increasing the yield of the low-impurity fraction of ferric hydroxide is a significant subject of further research. The Fe content of 52.2 wt. % makes such ferric hydroxide a valuable product for ironmaking after its conversion into hematite [94]. Moreover, such a product can be used optionally as an adsorbent for impurities not only in this processing route but also in other ones including waste water treatment processes [95,96]. After the removal of iron, other elements present in the leachate, such as Al, Zn, Mn, Mg, and Ca, can be recovered through sequential separation techniques involving pH adjustment [97], precipitation as carbonates [98] or sulfides [99], and related methods. The recovery of elements from leachates is an established industrial practice [100]. Identifying the most efficient strategy for the recovery of the remaining valuable components from hydrochloric acid leachates derived from ZKS remains a subject of ongoing investigation.

5. Conclusions

This study has demonstrated the feasibility of hydrochloric acid leaching for ZKS processing, followed by the selective recovery of copper and iron. The key contributions of this study can be summarized as follows:

• According to thermodynamic simulation followed by laboratory-scale experiments, HCl leaching proved to be the most effective in terms of reagent consumptions of 0.8 g HCl and 0.1 g H₂O₂ per gram of ZKS, with an L/S ratio of 5 mL/g, maintained at 70 °C for 180 min. Under these conditions, the recovery rates of Cu, Fe, and Zn reached 96.3%, 93.6%, and 76.8%, respectively.
• The remaining HCl leach residue was identified as predominantly graphite-based with 43.1 wt. % C, containing substantial fractions of non-crystalline silica, spinel phases, orthorhombic sulfur, and sodium-potassium halides. The remaining Cu, Fe, and Zn compounds in the residue were identified as various sulfides. After further treatment, the residue had the potential to be utilized in the Waelz process as a partial substitute for reducing agents.
• Cementation using iron powder enabled the recovery of 98.9% Cu and 91.2% As from the hydrochloric acid leachate.
• The two-step approach of iron precipitation achieved a recovery >90% of Fe, with about 70% concentrated in a low-impurity ferric hydroxide fraction with 52.2 wt. % Fe suitable for recycling.
• Thus, the integration of leaching and metal precipitation provides a scientifically validated and practically feasible approach based on a novel hydrochloric acid oxidative leaching route for the sustainable recycling of ZKS.