**1. Introduction**

In 2015, 193 governments adopted 17 sustainable development goals (SDGs) aiming to eradicate poverty, protect the planet, and ensure peaceful and prosperous lives for everyone by 2030 [1–3]. Number 13 of the SDGs—"Climate Action"—envisions drastic carbon dioxide (CO2) emission reductions to combat climate change and achieve a carbonneutral society by 2050 [4]. Copper (Cu), lead (Pb), and zinc (Zn) are classified as critical metals defined by the World Bank as essential in renewable energy and clean storage technologies (e.g., wind turbines, solar photovoltaic (PV) panels, batteries, etc.), and thus their demands are projected to significantly increase in the future to achieve a carbonneutral society [5,6].

Complex sulfide ores, some of the most important sources of critical metals, are typically composed of several metal sulfide minerals such as chalcopyrite (CuFeS2), galena (PbS), and sphalerite (ZnS). In the mineral processing of complex sulfide ores consisting of Cu-Pb-Zn sulfide minerals, flotation has been commonly adopted to produce concentrates of each mineral [7–16]. Some complex sulfide ores contain not only PbS but also anglesite (PbSO4), the presence of which is problematic for the separation of Cu, Pb, and Zn by flotation due to the unwanted activation of ZnS by Pb2+ released from PbSO<sup>4</sup> [17–19]. In the general flotation circuit of Cu-Pb-Zn sulfide ores, Cu and Pb sulfides are first recovered as froth, and then ZnS is recovered with the assistance of activators (e.g., CuSO4) [20,21]. When PbSO<sup>4</sup> is contained in complex sulfide ores, however, Pb2+ is readily released from PbSO<sup>4</sup> during conditioning and/or flotation due to its higher solubility compared to PbS the solubility product (Ksp) of PbSO<sup>4</sup> and PbS is 10−7.79 and 10−26.77, respectively [22]—and

40

activates the surface of ZnS via the formation of PbS-like compounds, as illustrated in Equation (1) [23,24];

$$\text{ZnS(s)} + \text{Pb}^{2+} = \text{PbS(s)} + \text{Zn}^{2+} \tag{1}$$

The activation of ZnS by Pb2+ is known to increase the ZnS floatability because of a higher affinity of PbS-like compounds for xanthate compared to ZnS [25–27]. As a result, ZnS is recovered as froth together with CuFeS<sup>2</sup> and PbS, making their separation difficult [17]. Therefore, a proper way of depressing ZnS floatability during flotation of complex sulfide ores containing PbSO<sup>4</sup> should be established.

In addition, PbSO<sup>4</sup> is typically distributed into flotation tailings because it has poor affinity for the collector (e.g., xanthate) [28,29]. The disposal of PbSO<sup>4</sup> into tailings storage facilities (TSFs) would cause lead pollution in the surrounding environment [30,31]. That is, the presence of PbSO<sup>4</sup> in complex sulfide ores would cause not only difficulties in the separation of ZnS from Cu and Pb sulfide minerals but also lead pollution in the surrounding environment of TSFs.

To address the above-mentioned problems, this study investigated the improvement of ZnS depression by extracting PbSO<sup>4</sup> and the recovery of extracted Pb2+ as zero-valent Pb by cementation using zero-valent iron (ZVI). Specifically, this study was devoted not only to optimizing the conditions for the extraction of PbSO<sup>4</sup> by ethylene diamine tetra acetic acid (EDTA) as well as the recovery of extracted Pb2+ as Pb<sup>0</sup> via cementation using ZVI but also to understanding how the absence/presence of PbSO<sup>4</sup> affected the floatability of ZnS through flotation experiments and surface characterizations of ZnS/PbSO<sup>4</sup> mixtures with and without EDTA pretreatment. Finally, a sustainable flowsheet for the processing of complex sulfide ores containing PbSO<sup>4</sup> with the combination of EDTA pretreatment, flotation, and cementation is proposed.

#### **2. Materials and Methods**

#### *2.1. Minerals and Reagents*

Five types of minerals were used in this study: sphalerite (ZnS, Kamioka Mine, Hida, Japan), lead sulfate (PbSO4, 98% purity), lead carbonate (PbCO3, 99% purity), lead chloride (PbCl2, 99% purity), and quartz (SiO2, 99% purity). The lead compounds and quartz were obtained from FUJIFILM Wako Pure Chemical Corporation (Osaka, Japan). A mineral specimen of ZnS was characterized using X-ray fluorescence spectroscopy (XRF, EDXL300, Rigaku Corporation, Tokyo, Japan) and X-ray powder diffraction (XRD, MultiFlex, Rigaku Corporation, Tokyo, Japan), and its chemical and mineralogical compositions ware shown in Table 1 and Figure 1, respectively. The XRD pattern of ZnS implied that ZnS was relatively pure due to the absence of the peaks of common minerals such as SiO2. In preparation for the flotation experiments, ZnS was ground by a vibratory disc mill (RS 200, Retsch Inc., Haan, Germany) and screened to obtain a size fraction of −75 + 38 µm.

**Table 1.** Chemical composition of a mineral specimen of ZnS based on XRF.


For the flotation experiments, potassium amyl xanthate (KAX, Tokyo Chemical Industry Co., Ltd., Tokyo, Japan) and methyl isobutyl carbinol (MIBC, FUJIFILM Wako Pure Chemical Corporation, Osaka, Japan) were used as collector and frother, respectively. Zinc sulfate (ZnSO4) and sodium sulfite (Na2SO3), purchased from FUJIFILM Wako Pure Chemical Corporation (Osaka, Japan), were used as depressants for ZnS. Sulfuric acid (H2SO4, FUJIFILM Wako Pure Chemical Corporation, Osaka, Japan) and sodium hydroxide (NaOH, FUJIFILM Wako Pure Chemical Corporation, Osaka, Japan) were used as pH adjusters. For the extraction experiments, ethylene diamine tetra acetic acid (EDTA, FUJIFILM Wako Pure Chemical Corporation, Osaka, Japan) was used as an extractant. For the cementation

experiments, zero-valent Fe (ZVI) powder (Fe<sup>0</sup> , −45 µm, FUJIFILM Wako Pure Chemical Corporation, Osaka, Japan) was used as a reductant. *Minerals* **2022**, *12*, x FOR PEER REVIEW 3 of 14

**Figure 1.** XRD pattern of a mineral specimen of ZnS. **Figure 1.** XRD pattern of a mineral specimen of ZnS.

#### *2.2. Experimental Methods*

*2.2. Experimental Methods*

For the flotation experiments, potassium amyl xanthate (KAX, Tokyo Chemical 2.2.1. Extraction Experiments of Lead Minerals Using EDTA

Industry Co., Ltd., Tokyo, Japan) and methyl isobutyl carbinol (MIBC, FUJIFILM Wako Pure Chemical Corporation, Osaka, Japan) were used as collector and frother, respectively. Zinc sulfate (ZnSO4) and sodium sulfite (Na2SO3), purchased from FUJIFILM Wako Pure Chemical Corporation (Osaka, Japan), were used as depressants for ZnS. Sulfuric acid (H2SO4, FUJIFILM Wako Pure Chemical Corporation, Osaka, Japan) and sodium hydroxide (NaOH, FUJIFILM Wako Pure Chemical Corporation, Osaka, Japan) were used as pH adjusters. For the extraction experiments, ethylene diamine tetra acetic acid (EDTA, FUJIFILM Wako Pure Chemical Corporation, Osaka, Japan) was used as an extractant. For the cementation experiments, zero-valent Fe (ZVI) powder (Fe<sup>0</sup> , −45 µm, FUJIFILM Wako Pure Chemical Corporation, Osaka, Japan) was used as a reductant. The extraction experiments of lead minerals were conducted in 50 mL Erlenmeyer flasks containing 1 g of the lead mineral and an EDTA solution (solid/liquid ratio: 1 g/10 mL). The solutions containing 100, 200, or 500 mM EDTA were used without pH adjustment (natural pH: 11.3). The flasks were shaken in a thermostat water bath shaker at a shaking speed of 120 rpm and an amplitude of 40 mm at 25 ◦C. After this, the leachates were collected by filtration using 0.2 µm syringe-driven membrane filters and immediately analyzed using an inductively coupled plasma atomic emission spectrometer (ICP-AES, ICPE 9820, Shimadzu Corporation, Kyoto, Japan) (margin of error = <sup>±</sup>2%) to measure the concentration of Pb2+ . The experiments were performed in triplicate to ascertain that the differences observed were statistically significant, and the extraction efficiency of the lead mineral (E*Pb*) was calculated using the following equation:

$$\mathbf{E}\_{Pb} = \frac{[\text{Pb}]\_i}{[\text{Pb}]\_{tot}} \times 100\tag{2}$$

The extraction experiments of lead minerals were conducted in 50 mL Erlenmeyer where [Pb] *tot* and [Pb] *i* are total Pb (mM) and extracted Pb (mM), respectively.

flasks containing 1 g of the lead mineral and an EDTA solution (solid/liquid ratio: 1 g/10 mL). The solutions containing 100, 200, or 500 mM EDTA were used without pH 2.2.2. Cementation Experiments of the Extracted Pb2+

adjustment (natural pH: 11.3). The flasks were shaken in a thermostat water bath shaker at a shaking speed of 120 rpm and an amplitude of 40 mm at 25 °C. After this, the leachates were collected by filtration using 0.2 μm syringe-driven membrane filters and immediately analyzed using an inductively coupled plasma atomic emission spectrometer (ICP-AES, ICPE 9820, Shimadzu Corporation, Kyoto, Japan) (margin of error = ±2%) to measure the concentration of Pb2+. The experiments were performed in triplicate to ascertain that the differences observed were statistically significant, and the extraction efficiency of the lead mineral (E) was calculated using the following equation: E = [Pb] [Pb] × 100 (2) where [Pb] and [Pb] are total Pb (mM) and extracted Pb (mM), respectively. 2.2.2. Cementation Experiments of the Extracted Pb2+ The cementation experiments of Pb2+ extracted from PbSO<sup>4</sup> were conducted in 50 mL Erlenmeyer flasks. A volume of 10 mL of the leachate collected by filtration using 0.2 μm The cementation experiments of Pb2+ extracted from PbSO<sup>4</sup> were conducted in 50 mL Erlenmeyer flasks. A volume of 10 mL of the leachate collected by filtration using 0.2 µm syringe-driven membrane filters after the extraction experiments ([EDTA], 500 mM; S/L ratio, 1 g/10 mL; time, 30 min) was added to the flask, and ultrapure nitrogen gas (N2; 99.99%) was introduced for 15 min to remove the dissolved oxygen (DO) present in the leachate. Then, a known amount of ZVI powder was added to the flask, and N<sup>2</sup> was further introduced for 5 min. The flask was tightly capped with a silicon rubber plug and parafilm and shaken in a thermostat water bath shaker at a shaking speed of 120 rpm and an amplitude of 40 mm at 25 ◦C. After this, the leachate was filtered using a 0.2 µm syringe-driven membrane filter, and the filtrate was immediately analyzed using ICP-AES to measure the concentration of Pb2+. Meanwhile, the residue was washed thoroughly with distilled water, dried in a vacuum-drying oven at 40 ◦C, and analyzed by scanning electron microscopy with energy-dispersive X-ray spectroscopy (SEM-EDS, JSM-IT200TM, JEOL Co., Ltd., Tokyo, Japan). The experiments were performed in triplicate to ascertain that the differences observed were statistically significant. Pb recovery (R*Pb*) was calculated using the following equation:

99.99%) was introduced for 15 min to remove the dissolved oxygen (DO) present in the leachate. Then, a known amount of ZVI powder was added to the flask, and N<sup>2</sup> was further introduced for 5 min. The flask was tightly capped with a silicon rubber plug and parafilm and shaken in a thermostat water bath shaker at a shaking speed of 120 rpm and

$$\mathbf{R}\_{Pb} = \frac{[\mathbf{P}\mathbf{b}]\_I - [\mathbf{P}\mathbf{b}]\_F}{[\mathbf{P}\mathbf{b}]\_I} \times 100\tag{3}$$

where [Pb] *I* and [Pb]*<sup>F</sup>* are the initial and the final concentrations of Pb (mM) in the leachate, respectively.

#### 2.2.3. Flotation Experiments

The flotation experiments were carried out using an agitator-type flotation machine (ASH-F30H, Kankyo-kanri Engineering, Akita, Japan) equipped with a 400 mL flotation cell under the following conditions: pH, 6.5; temperature, 25 ◦C; pulp density, 5%; impeller speed, 1000 rpm; air flow rate, 1 L/min. In a 500 mL beaker, a model sample containing 15 g of ZnS and 5 g of PbSO<sup>4</sup> or SiO<sup>2</sup> was suspended in 300 mL of distilled water, and the supernatant was decanted to remove fine particles (<38 µm) [17]. After desliming, a model sample was repulped to 400 mL with distilled water in the flotation cell and conditioned for 3 min after adding the following reagents in sequence: ~5 kg/t of ZnSO<sup>4</sup> (100 ppm Zn2+), 1 kg/t of Na2SO3, 20 g/t of KAX, and 20 µL/L of MIBC. Afterwards, air was introduced at a flowrate of 1 L/min, and froth was recovered for 3 min. The recovered froth/tailing products were dried at 105 ◦C for 24 h and analyzed by XRF to determine the recovery of Zn. The experiments were performed in duplicate to ascertain that the differences observed were statistically significant. An aliquot of the pulp of about 5 mL was collected 3 min after adding the depressant, filtered through a 0.2 µm syringe-driven membrane filter, and immediately analyzed using ICP-AES to check the extent of lead activation of ZnS. Meanwhile, the collected residues were washed thoroughly with distilled water, dried in a vacuum-drying oven at 40 ◦C, and analyzed by X-ray photoelectron spectroscopy (XPS, JPS-9200, JEOL Co., Ltd., Tokyo, Japan). The XPS analysis was conducted using an Al Kα X-ray source (1486.7 eV) operated at 100 W (voltage = 10 kV; current = 10 mA) under ultrahigh vacuum conditions (~10−<sup>7</sup> Pa). The binding energies of photoelectrons were calibrated using C1s (285 eV) or Zn2p3/2 (1022.0 eV) for charge correction. The XPS data were analyzed by Casa XPS, and deconvolutions of the spectra were carried out using an 80% Gaussian–20% Lorentzian peak model and a Shirley background [32–35].

In addition, flotation experiments of ZnS/PbSO<sup>4</sup> mixture after EDTA pretreatment were conducted to check how effective the removal of PbSO<sup>4</sup> was on the depression of ZnS floatability. A mixture of 15 g ZnS and 5 g PbSO<sup>4</sup> was mixed with 200 mL of a 500 mM EDTA solution (i.e., solid/liquid ratio: 20 g/200 mL) and shaken in a thermostat water bath shaker at a shaking speed of 120 rpm and an amplitude of 40 mm at 25 ◦C for 30 min. After this, the leachate was filtered using a 0.2 µm syringe-driven membrane filter, the filtrate was immediately analyzed using ICP-AES to measure the concentration of Pb2+, and the cementation experiments to recover the extracted Pb2+ in the filtrate were carried out using 1 g/10 mL of ZVI for 24 h under the same conditions and with the same procedures mentioned in Section 2.2.2. Meanwhile, the residue obtained after EDTA pretreatment was washed thoroughly with distilled water and deslimed to remove fine particles [17]. After desliming, the residue was repulped to 400 mL with distilled water, and then flotation was conducted under the same conditions mentioned above. An aliquot of the pulp of about 5 mL was also collected 3 min after adding the depressant (ZnSO4) and filtered through a 0.2 µm syringe-driven membrane filter, and the filtrate and residue were analyzed by ICP-AES and XPS, respectively.

#### **3. Results and Discussion**
