Flotation Kinetics of Oxidized Lead–Zinc Ore in the Eh–pH System Using Calcium Polysulfide

Abstract

The flotation of oxidized lead–zinc ores is challenging due to the low floatability of oxidized minerals and their weak interaction with conventional reagents. This study investigates the influence of pulp electrochemical parameters, namely redox potential (Eh) and pH, on the flotation kinetics of oxidized lead–zinc ore from the Koskuduk deposit. The results showed that sodium sulfide selectively activates lead-bearing minerals, increasing Pb recovery to 40.74%, while Zn recovery remains low at approximately 12%. In contrast, the polysulfide–lime system S:CaO:H₂O provides more uniform and stable sulfidization of oxidized mineral surfaces, increasing recovery to 65.10% for Pb and 56.89% for Zn. The highest recoveries were achieved within an Eh range of −120 to −180 mV at pH 11–12. Kinetic studies demonstrated that the main contribution to metal recovery occurs during the first 2–6 min of flotation. These results indicate that flotation efficiency is controlled not only by reagent type but also by the electrochemical state of the pulp, confirming that calcium polysulfide is a promising alternative sulfidizing reagent for processing oxidized lead–zinc ores.

1. Introduction

The processing of oxidized lead–zinc ores remains one of the most difficult problems in modern mineral processing. With the gradual depletion of easily beneficiated sulfide deposits, increasing attention is being paid to refractory oxidized and mixed ores. These ores usually contain lead and zinc in the form of carbonates, sulfates, silicates, hydroxides, and finely dispersed secondary minerals. Compared with sulfide minerals, oxidized lead–zinc minerals are characterized by higher surface hydrophilicity, weaker interaction with sulfhydryl collectors, complex associations with gangue minerals, and lower natural floatability. As a result, direct flotation of such ores is generally inefficient, and preliminary surface sulfidization is required to improve collector adsorption and flotation recovery [1,2,3,4].

Previous studies have shown that the flotation behavior of oxidized lead–zinc ores is strongly affected by mineralogical composition, degree of liberation, surface chemistry, pulp chemistry, and electrochemical conditions. In particular, the formation of sulfide-like films on oxidized mineral surfaces and their interaction with collectors are closely related to the pH and redox potential (Eh) of the pulp [5,6,7,8]. These factors are especially important for ores containing finely disseminated oxidized lead and zinc minerals, iron hydroxides, clay minerals, and a high proportion of gangue phases.

Similar difficulties are characteristic of polymetallic ores from Kazakhstan, including the Shalkiya, Shaimerden, and Koskuduk deposits. These ores are often characterized by a high degree of oxidation, a low proportion of primary sulfide minerals, and the predominance of secondary oxygen-containing forms of lead and zinc. Previous studies have shown that the beneficiation of such ores is complicated by low mineral reactivity, fine dissemination, and strong dependence of flotation performance on reagent scheme and pulp electrochemical conditions [9,10,11]. Therefore, the Koskuduk oxidized lead–zinc ore used in this study represents a typical difficult-to-process object requiring optimization of both reagent regime and Eh–pH conditions.

Several approaches have been used for the processing of oxidized lead–zinc ores, including pyrometallurgical, hydrometallurgical, and flotation methods [12,13,14]. Pyrometallurgical processes are usually associated with high energy consumption and environmental limitations, whereas hydrometallurgical treatment may require high reagent consumption and can be difficult to apply to complex low-grade ores. Therefore, sulfidization flotation remains one of the most widely used and promising methods for the pre-concentration of oxidized lead–zinc ores. In this process, the surface of oxidized minerals is converted into a sulfide-like layer, which promotes the adsorption of sulfhydryl collectors and improves the floatability of valuable minerals [15,16,17].

Sodium sulfide (Na₂S) is the most commonly used sulfidizing reagent in the flotation of oxidized lead–zinc ores. However, its action is often selective and difficult to control. Na₂S rapidly releases sulfide species into the pulp and effectively sulfidizes lead-bearing minerals, especially cerussite, through the formation of a PbS surface film. This usually improves lead recovery, but the sulfidization of zinc-bearing oxidized minerals is less effective because the ZnS-like surface layer is less stable and may require additional activation [18,19,20,21]. In addition, excessive sulfide ions may form unstable colloidal products and reduce flotation selectivity. Under less alkaline conditions, free sulfide species may also contribute to the formation of toxic hydrogen sulfide gas, which creates environmental and operational concerns [18,20].

In this regard, polysulfide reagents are of interest as alternative sulfidizing agents. Calcium polysulfide, also known as lime sulfur, is generally represented as CaSx, where x denotes the number of sulfur atoms in the polysulfide chain [22]. In the present study, the notation S:CaO:H₂O refers to the preparation ratio of sulfur, lime, and water used to obtain the calcium polysulfide solution, rather than to the final molecular formula of the reagent. Compared with Na₂S, calcium polysulfide contains sulfur mainly in polysulfide forms (S_x²⁻), which can participate in more gradual surface reactions and promote more uniform sulfidization of oxidized mineral surfaces. Since calcium polysulfide is prepared in a strongly alkaline lime-containing medium and sulfur is partly present in polysulfide chains, the release of free sulfide species is more gradual than in the case of Na₂S. Therefore, under controlled alkaline flotation conditions, the tendency for H₂S formation may be reduced [18,20,22].

The oxidized lead–zinc ore from the Koskuduk deposit contains a large proportion of quartz and micaceous gangue minerals, while lead and zinc are mainly present as oxidized, finely dispersed, and poorly liberated forms associated with iron hydroxides and clay minerals. Such mineralogical characteristics limit the efficiency of direct flotation and justify the need to optimize sulfidization conditions through Eh–pH regulation. Although recent studies have addressed the process mineralogy, reagent schemes, and flotation behavior of oxidized lead–zinc ores [3,4,5,10,11], the combined effect of calcium polysulfide sulfidization, Eh–pH regulation, and flotation kinetics for the Koskuduk oxidized lead–zinc ore has not been systematically investigated. To the best of our knowledge, this work provides the first integrated Eh–pH mapping and kinetic assessment of calcium polysulfide-assisted flotation for this ore type. This approach makes it possible to evaluate not only the final recovery values, but also the relationship between reagent action, pulp electrochemical state, and the time-dependent recovery of Pb and Zn. Therefore, the aim of the present study is to investigate the influence of pulp electrochemical parameters (Eh and pH) on the flotation kinetics of oxidized lead–zinc ore from the Koskuduk deposit and to evaluate the effectiveness of a calcium polysulfide–lime system in comparison with conventional sodium sulfide. Special attention is paid to the relationship between reagent type, Eh–pH conditions, flotation kinetics, and the recovery of lead and zinc.

2. Materials and Methods

In the experiments, a sample of oxidized lead–zinc ore from the Koskuduk deposit was used.

The mineralogical composition of the ore was investigated by reflected-light optical microscopy on polished sections. The analysis was carried out using an Olympus BX53 metallurgical microscope (Olympus Corporation, Hachioji, Tokyo, Japan) equipped with a SIAMS XS-3CU digital video camera (SIAMS, Yekaterinburg, Russia). Image processing and quantitative analysis were performed using the Mineral C7 software package, version 7 (SIAMS, Yekaterinburg, Russia).

The relative contents of major mineral groups were estimated by quantitative image analysis of polished sections using the Mineral C7 software. The calculation was based on the areal proportion of identified mineral phases in representative microscopic fields, followed by normalization to 100%. The total content of quartz and micaceous minerals was determined as the sum of the corresponding gangue mineral phases identified during optical-mineralogical analysis.

The studied material had a particle size of −2.0 + 0 mm. It was established that the ore is characterized by a predominance of gangue minerals, represented by quartz and micaceous minerals (albite, muscovite, and chlorite), whose total content exceeds 80%. The ore mineralization is unevenly distributed and is mainly represented by secondary oxidized forms of lead, zinc, and iron, with a subordinate role of primary sulfides.

Galena (PbS) is present in minor amounts in the form of relict grains measuring 20–150 µm, characterized by intense corrosion and partial replacement by secondary lead minerals.

The principal form of lead occurrence is cerussite (PbCO₃), which develops after galena and forms fine-dispersed porous aggregates measuring 5–50 µm, closely associated with iron hydroxides and clay minerals.

Anglesite (PbSO₄) is identified in subordinate amounts and develops together with cerussite after galena. It forms thin rims, veinlets, and micro-scale replacement zones measuring <1-5-30 μm. Anglesite is often localized along fractures and grain boundaries of galena, indicating its formation at a later stage of supergene processes.

Sphalerite (ZnS) occurs as isolated relict grains measuring 3–70 μm, strongly corroded and almost completely destroyed during supergene processes. The corrosion replacement of galena by cerussite and the inclusions of pyrite are shown in Figure 1.

The development of anglesite together with cerussite after galena and the corrosion of sphalerite in association with galena, covellite, and pyrite are shown in Figure 2 and Figure 3.

The bulk of zinc is represented by finely dispersed oxidized forms that do not form independent mineral phases and are distributed within a clay-iron matrix.

The chemical composition of the original ore was determined by inductively coupled plasma optical emission spectrometry (ICP-OES) using an Agilent 725ES spectrometer (Agilent Technologies, Santa Clara, CA, USA) after preliminary acid digestion of the samples.

According to ICP-OES data, the chemical composition of the studied ore is characterized by the following major components (wt.%): SiO₂—32.39; Al₂O₃—7.09; K₂O—4.39; CaO—1.26; TiO₂—0.72; MnO—0.35; P₂O₅—0.46.

The content of valuable components is (wt.%): Zn—0.56; Pb—0.58; Cu—0.023; Fe—9.02. The sulfur content is 0.28 wt.%. Trace elements are present in minor amounts (wt.%): As—0.014; Mo—0.002; Bi—0.003; Cr—0.008; Ni—0.002; Sn—0.002; V—0.01; W—0.021; Ba—0.042; Sr—0.021; Rb—0.030.

The phase chemical analysis of the studied ore was carried out using sequential selective leaching, which allows determination of the forms of occurrence of lead, zinc, and iron depending on their chemical stability. The method is based on stepwise extraction of metal compounds using reagents with different dissolving capacities, enabling the separation of water-soluble, oxidized, sulfide, and refractory forms.

For iron, the Fe²⁺ and Fe³⁺ forms were additionally determined, allowing the degree of ore oxidation to be assessed. Element concentrations in the solutions were measured by ICP-OES, and the relative proportions of the forms were calculated as percentages of the total elemental content. The results of the chemical phase analysis of the oxidized ore from the Koskuduk deposit are presented in

Table 1. Results of the chemical phase analysis of the oxidized ore from the Koskuduk deposit.

Name of Components	Content
	Absolute, %	Relative, %
Forms of zinc occurrence
− Oxidized	0.21	37.53
− Sulfides	0.07	12.61
− Refractory (poorly soluble)	0.28	49.86
Total	0.56	100.0
Forms of lead occurrence
− Oxygen-containing	0.39	66.58
− Galena	0.09	16.05
− Residual forms	0.10	17.37
Total	0.58	100.0
Forms of iron occurrence
− Sulfides	<0.04	-
− Ferrous iron Fe2+	4.49	49.77
− Ferric iron Fe3+	4.53	50.23
Total	9.02	100.0

.

The obtained results indicate a high degree of ore oxidation and the predominance of finely dispersed and sorbed forms of lead and zinc, closely associated with iron hydroxides and clay minerals. Such modes of occurrence significantly reduce the degree of mineralogical liberation and complicate the recovery of valuable components by direct flotation.

The low proportion of sulfide forms explains the limited effectiveness of conventional flotation methods and justifies the need for sulfidization treatment, as well as optimization of pulp electrochemical parameters, including pH and redox potential (Eh).

Reagents. The reagents were added sequentially with prior pulp conditioning. The total conditioning time was 3 min. In the first stage, a sulfidizing agent (S:CaO:H₂O) was introduced into the pulp in order to form a sulfide film on the surface of oxidized minerals. Here, S:CaO:H₂O denotes the mass ratio of sulfur, lime, and water used for the preparation of the calcium polysulfide solution; the active polysulfide reagent is generally represented as CaSx. This was followed by the addition of a collector-potassium butyl xanthate (Kx, 100 g/t)-which ensured hydrophobization of the activated surface. In the final stage, the frother T-92 was added at a dosage of 10 g/t.

The sample was ground in a laboratory ball mill MShL-7 (Mechanobr-Technica, Saint Petersburg, Russia). A 1 kg sample with a particle size of −2 + 0 mm was used as the feed material. Grinding was carried out under wet conditions at a solids content of 60% in the pulp, which corresponds to industrial processing conditions.

The grinding time was 10 min. This value was determined based on preliminary grindability tests of the ore and ensured the required fineness of grinding-68.54% of the −0.071 mm size fraction.

Flotation tests were carried out on a laboratory flotation machine VEKTiS with a 3 L cell volume. The pulp was prepared at a solids content of 30% by mass (ore specific gravity: 2.70 g/cm³).

The aeration unit of the flotation machine is of the pneumatic–mechanical type. Pulp mixing was carried out using an impeller rotating at a speed of 1700 rpm. The froth product was removed using a froth scraper at a rotation speed of 3 rps.

Air was supplied to the flotation zone at a pressure of not less than 5 bar. The air flow rate was up to 10 NL/min, while the specific air consumption was at least 0.8 m³/(m²·min), which ensured the formation of a stable froth layer.

Flotation tests were carried out at a pulp temperature of 26 °C.

Control of pH and redox potential (Eh, ORP) during laboratory flotation tests was carried out using a potentiometric method. A HI 1230 combined electrode (Hanna Instruments, Woonsocket, Rhode Island, USA; pH range 0–14) was used for pH measurement, while a HI 3131 electrode (Hanna Instruments, Woonsocket, RI, USA; ORP range ±399.9 mV) was used for ORP determination. ORP values were recorded after reagent addition and prior to the start of pulp aeration.

3. Results

Taking into account the mineralogical composition of the studied ore and the predominance of oxidized forms of lead and zinc, a laboratory flotation flowsheet was developed, including grinding, main flotation, and scavenger flotation stages (shown in Figure 4). The use of a two-stage scheme is justified by the need to improve the recovery of finely dispersed and refractory particles. This is supported by a number of studies, which demonstrate that flotation of oxidized lead–zinc ores requires prior sulfidization and multi-stage circuits to enhance the recovery of valuable components [15,16,17].

Comparison of the flotation results presented in

Table 2. Qualitative and quantitative indicators of the main flotation concentrates depending on the sulfidizing agent dosage.

Dosage of S:CaO:H2O, g/kg	Product	Yield, %	Grade, %		Recovery, %
			Zn	Pb	Zn	Pb
Base experiment
0	Concentrate 1	3.18	0.69	0.98	3.90	5.39
	Concentrate 2	2.80	1.69	2.46	8.45	11.91
	∑ (Total)	5.98	1.16	1.68	12.35	17.30
	Tailings	94.02	0.52	0.51	87.65	82.70
	Feed (initial sample)	100.0	0.56	0.58	100.0	100.0
Test 1
Na2S—500 g/t	Concentrate 1	2.89	1.50	7.18	7.75	35.73
	Concentrate 2	3.65	0.78	0.80	5.11	5.01
	∑ (Total)	6.54	1.10	3.61	12.85	40.74
	Tailings	93.46	0.52	0.37	87.15	59.26
	Feed (initial sample)	100.0	0.56	0.58	100.0	100.0
Test 2
1:0.5:8.5	Concentrate 1	5.43	4.74	4.27	45.94	40.02
	Concentrate 2	3.43	0.79	0.96	4.82	5.69
	∑ (Total)	8.86	3.21	2.99	50.76	45.71
	Tailings	91.14	0.30	0.35	49.24	54.29
	Feed (initial sample)	100.0	0.56	0.58	100.0	100.0
Test 3
2:1:17	Concentrate 1	9.70	2.87	3.61	49.66	60.39
	Concentrate 2	3.33	0.62	1.64	3.71	9.41
	∑ (Total)	13.03	2.29	3.11	53.37	69.79
	Tailings	86.97	0.30	0.20	46.63	30.21
	Feed (initial sample)	100.0	0.56	0.58	100.0	100.0
Test 4
3:1.5:25.5	Concentrate 1	12.14	2.34	2.84	50.82	59.41
	Concentrate 2	4.55	0.75	0.72	6.07	5.68
	∑ (Total)	16.69	1.91	2.26	56.89	65.10
	Tailings	83.31	0.29	0.24	43.11	34.90
	Feed (initial sample)	100.0	0.56	0.58	100.0	100.0
Test 5
4:2:34	Concentrate 1	7.88	3.38	4.16	47.60	56.52
	Concentrate 2	5.48	0.62	0.69	6.04	6.50
	∑ (Total)	13.36	2.25	2.74	53.64	63.03
	Tailings	86.64	0.30	0.25	46.36	36.97
	Feed (initial sample)	100.0	0.56	0.58	100.0	100.0
Test 6
5:2.5:42.5	Concentrate 1	8.78	2.99	3.54	46.91	53.68
	Concentrate 2	8.43	0.58	0.60	8.74	8.76
	∑ (Total)	17.22	1.81	2.10	55.64	62.45
	Tailings	82.78	0.30	0.26	44.36	37.55
	Feed (initial sample)	100.0	0.56	0.58	100.0	100.0
Test 7
6:3:51	Concentrate 1	15.29	1.45	1.82	39.57	47.90
	Concentrate 2	4.55	0.67	0.70	5.48	5.50
	∑ (Total)	19.84	1.27	1.56	45.05	53.39
	Tailings	80.16	0.38	0.34	54.95	46.61
	Feed (initial sample)	100.0	0.56	0.58	100.0	100.0

shows that the nature of the sulfidizing agent has a determining influence on the recovery of valuable components from the oxidized lead–zinc ore of the Koskuduk deposit. In the base experiment without the use of a sulfidizing agent, the total recovery of Zn and Pb was 12.35% and 17.30%, respectively, confirming the low efficiency of direct flotation of the studied ore.

When Na₂S is used at a dosage of 500 g/t, lead recovery increases to 40.74%, whereas zinc recovery remains almost unchanged at 12.82%. This indicates that Na₂S predominantly activates lead-bearing minerals, primarily cerussite and other oxygen-containing lead phases, through the formation of a PbS film, which promotes xanthate adsorption [18]. At the same time, the effect of Na₂S on zinc-bearing minerals is limited, since the formation of ZnS does not always lead to improved floatability without additional activation [19]. The effect of sulfidizer type and dosage on metal recovery is shown in Figure 5.

A different trend is observed when using the polysulfide–lime reagent S:CaO:H₂O. In all experiments with this reagent, the recovery of both zinc and lead increases significantly compared not only with the base experiment but also with the case of Na₂S application. Thus, at a ratio of 1:0.5:8.5, Zn and Pb recoveries reach 50.76% and 45.71%, respectively, while at a ratio of 2:1:17 they increase to 53.37% and 69.79%, respectively. The maximum zinc recovery of 56.89% is obtained at a ratio of 3:1.5:25.5, whereas lead recovery under these conditions is 65.10%.

To provide a clearer representation of the combined effect of electrochemical parameters, a flotation regime map (Eh–pH diagram) was constructed (Figure 6). The diagram illustrates the relationship between redox potential, pH of the medium, and the recovery of zinc and lead.

The results showed that the highest recovery values are achieved within a specific electrochemical range corresponding to Eh values from −120 to −180 mV and pH 11–12. This region is characterized as optimal for the sulfidization of oxidized mineral surfaces and subsequent collector adsorption.

Outside this region, a significant decrease in recovery is observed. In particular, under strongly reducing conditions (Eh ≈ −320 mV), corresponding to the use of Na₂S, a marked increase in lead recovery is observed, while zinc recovery remains low. This confirms the selective action of Na₂S toward lead-bearing minerals.

This result is consistent with literature data, according to which Na₂S effectively sulfidizes the surface of oxidized lead minerals with the formation of a PbS film, promoting xanthate adsorption and increasing lead recovery [20]. At the same time, the efficiency of sulfidization for zinc-bearing minerals is significantly lower, since their surface is characterized by a smaller number of active sites and requires additional activation by heavy metal ions [21].

Kinetics of Flotation

Flotation kinetic studies were carried out to evaluate the effect of process time on the recovery of valuable components. Concentrate was collected in fractions at 2–3 min intervals, which made it possible to monitor the dynamics of metal recovery over time. The qualitative and quantitative indicators of the bulk concentrate depending on the main flotation time are presented in

Table 3. Qualitative and quantitative indicators of the bulk concentrate depending on the main flotation time.

Product	Yield, %	Content, %		Recovery, %
		Zn	Pb	Zn	Pb
1st fraction	3.85	3.96	4.38	27.22	29.06
2nd fraction	2.56	3.34	4.24	15.29	18.71
3rd fraction	2.66	1.91	2.49	9.07	11.40
4th fraction	2.45	1.30	2.11	5.68	8.89
5th fraction	1.92	0.87	0.94	2.98	3.12
6th fraction	2.50	0.58	0.50	2.60	2.14
7th fraction	1.48	0.57	0.79	1.50	2.01
8th fraction	1.87	0.45	0.64	1.49	2.06
9th fraction	2.70	0.28	0.31	1.35	1.43
10th fraction	0.76	0.58	0.83	0.79	1.09
∑1–10 fractions	22.75	1.67	2.04	67.97	79.91
Tailings	77.26	0.23	0.15	32.03	20.09
Feed (initial sample)	100.0	0.56	0.58	100.0	100.0

.

To quantitatively describe the flotation kinetics, the cumulative recovery data were fitted using the classical first-order flotation kinetic model [23]:

$$R_t= R_{\infty }(1-e^{-kt})$$

where R_t is the cumulative recovery at flotation time t, R_∞ is the ultimate recovery, and k is the flotation rate constant. The cumulative recoveries were calculated by summing the recovery values of individual concentrate fractions. The fitting results are presented in

Table 4. First-order flotation kinetic parameters for Zn and Pb recovery.

Component	R∞, %	k, min−1	R2
Zn	65.76	0.21	0.97
Pb	77.95	0.19	0.99

.

The high R² values indicate that the first-order kinetic model adequately describes the flotation behavior of both Zn and Pb. The rate constant for Zn (0.208 min⁻¹) is slightly higher than that for Pb (0.191 min⁻¹), indicating a somewhat faster initial flotation response of zinc-bearing particles under the studied conditions. However, the ultimate recovery of Pb (77.95%) is higher than that of Zn (65.76%), which is consistent with the higher cumulative Pb recovery observed in the flotation fractions. The main recovery of both metals occurs during the first 2–8 min, corresponding to the most favorable Eh–pH region, whereas the later decrease in recovery rate is associated with the shift in Eh toward less negative values and a slight decrease in pH. The changes in electrochemical parameters during flotation over time are shown in Figure 7.

Figure 7 shows the trajectory of changes in the redox potential and pulp pH during flotation. It was established that at the initial stage of the process (2–6 min), Eh values range from −175 to −157 mV at pH 11.6–11.8, which corresponds to optimal flotation conditions.

As flotation time increases, a shift in Eh toward less negative values (up to −120 mV) is observed, accompanied by a slight decrease in pH, which is accompanied by a reduction in the recovery rate of valuable components.

4. Discussion

The sulfidization of oxidized lead and zinc mineral surfaces is a heterogeneous process involving chemical and electrochemical reactions at the phase interface. For lead minerals, in particular cerussite (PbCO₃), the process can be described by the following reaction [18]:

PbCO₃ + S²⁻ → PbS + CO₃²⁻

The formation of a PbS sulfide film ensures hydrophobization of the surface and promotes the adsorption of sulfhydryl collectors. As shown in the work of Li et al. [19], this process proceeds with the formation of a core–shell PbCO₃/PbS structure, which significantly enhances the floatability of the mineral.

For zinc-bearing minerals, such as smithsonite (ZnCO₃), the sulfidization process proceeds less efficiently and can be described by the following reaction [19,21]:

ZnCO₃ + S²⁻ → ZnS + CO₃²⁻

However, as noted in the work of Jia et al. [21], the resulting ZnS film is less stable and requires additional activation, which explains the low zinc recovery when Na₂S is used.

The use of calcium polysulfide (CaSx) alters the sulfidization mechanism due to the involvement of polysulfide ions (Sx²⁻), which are capable of forming more stable surface compounds [24]:

PbCO₃ + S_x²⁻ → PbS + S_x−1²⁻ + CO₃²⁻

ZnCO₃ + S_x²⁻ → ZnS + S_x−1²⁻ + CO₃²⁻

In contrast to Na₂S, polysulfide systems provide a more uniform coating of mineral surfaces and reduce the likelihood of forming colloidal reaction products. In addition, as shown in the works of Peng et al. [6] and Liu et al. [24], such systems contribute to the stabilization of the redox potential (Eh), creating favorable conditions for the formation of a stable sulfide film.

Compared with sodium sulfide, calcium polysulfide changes the sulfidization mechanism due to the presence of polysulfide ions (Sx²⁻), which are involved in stepwise surface reactions and provide a more gradual formation of sulfide-like layers on oxidized mineral surfaces. When Na₂S is used, sulfide ions are rapidly released into the pulp and mainly promote the sulfidization of lead-bearing oxidized minerals, especially cerussite, through the formation of a PbS film. This mechanism explains the selective increase in Pb recovery, whereas Zn recovery remains limited because the sulfidized layer formed on zinc-bearing oxidized minerals is less stable and often requires additional activation [18,19,20,21].

In contrast, calcium polysulfide supplies polysulfide species that can promote more uniform sulfidization of both lead- and zinc-bearing oxidized minerals and contribute to the stabilization of the pulp redox potential. Such electrochemical stabilization is important because the formation of surface sulfide films and collector adsorption are strongly controlled by Eh and pH conditions [6,24]. Therefore, the increased recovery of lead and zinc when using calcium polysulfide is attributed to the combined effect of more efficient surface sulfidization, more favorable collector adsorption, and stabilization of the pulp electrochemical regime.

A direct comparison of Pb and Zn recoveries with previously published results is difficult because flotation performance strongly depends on ore grade, mineralogical composition, degree of oxidation, liberation size, reagent scheme, pulp chemistry, and flotation circuit configuration. Therefore, the comparison presented in

Table 5. Comparison of flotation performance reported for oxidized lead–zinc ores and minerals.

Study	Object	Reagent/Process	Recovery Level	Comment
Önal et al. [15]	Oxide Pb–Zn ore	Sulfidization flotation	High Pb recovery; Zn mostly remained problematic	Results depend on multi-stage flotation and ore type.
Motovilov et al. [10]	Oxidized Pb–Zn ore	Gravity–flotation scheme	Pb recovery above 70%	Combined beneficiation scheme was used.
Jia et al. [21]	Hemimorphite	Na2S + Pb(II) activation	High Zn recovery for single mineral	Not directly comparable with complex ore.
This study	Koskuduk oxidized Pb–Zn ore	Calcium polysulfide–lime system, Eh–pH control	Pb 65.10%, Zn 56.89%	Balanced recovery of both Pb and Zn from complex oxidized ore.

should be considered as an indicative assessment rather than a direct ranking of process efficiency.

As shown in Table 5, previous studies reported high lead recovery when multi-stage sulfidization flotation or combined gravity–flotation schemes were applied. High zinc recovery was also reported for individual zinc oxide minerals such as hemimorphite under specific activation conditions. However, these results are not directly equivalent to the present study because the Koskuduk ore is a low-grade, highly oxidized Pb–Zn ore with finely dispersed valuable components and a high proportion of gangue minerals. Under these conditions, the obtained Pb recovery of 65.10% and Zn recovery of 56.89% indicate that the calcium polysulfide–lime system provides a competitive and balanced recovery of both metals, rather than selective improvement of only one component.

However, for industrial application, several practical aspects should be considered. The efficiency of calcium polysulfide may depend on the stability of the reagent solution, dosage accuracy, and the control of pulp Eh–pH conditions. In addition, variations in ore composition, pulp chemistry, and flotation residence time may influence the reproducibility of the sulfidization process. Therefore, pilot-scale tests are recommended to optimize reagent consumption and confirm process stability under plant conditions [6,24].

From a practical perspective, Na₂S is a widely available and relatively simple reagent; however, its rapid release of sulfide species requires strict pH control to minimize H₂S formation [18,20]. Calcium polysulfide may reduce this risk under alkaline conditions because sulfur is partly present in polysulfide chains and is released more gradually [22]. At the same time, its industrial use requires control of reagent preparation, solution stability, dosage, and Eh–pH monitoring. Therefore, the economic feasibility of the calcium polysulfide–lime system should be evaluated in pilot-scale tests by comparing reagent cost, recovery improvement, and process stability.

5. Conclusions

It was established that the use of sodium sulfide (Na₂S) in the flotation of oxidized lead–zinc ore results in the selective activation of lead (up to 40.74% recovery) with low zinc recovery, which limits its application for the processing of complex ores.

It was shown that the use of the polysulfide–lime system S:CaO:H₂O provides more efficient and uniform sulfidization of mineral surfaces, which makes it possible to increase recovery to 65.10% for Pb and 56.89% for Zn.

It was determined that the optimal flotation conditions correspond to the Eh range of −120 to −180 mV and pH 11–12, at which the maximum degree of sulfidization and collector adsorption is achieved.

It was established that flotation kinetics are governed by the dynamics of changes in the electrochemical state of the pulp, with the main contribution to metal recovery occurring during the first 2–6 min of the process.

The obtained results confirm that the use of polysulfide systems and control of Eh–pH parameters is an effective approach for improving metal recovery from oxidized lead–zinc ores.