Process Mineralogy Study and Flotation Testwork of a Complex Lead–Gold Rougher Concentrate

Abstract

A lead–gold rougher concentrate was studied to investigate the efficiency of mineral processing. Using process mineralogy as the guiding theory, mineralogical parameters such as chemical composition, mineral composition, mineral particle size, and symbiotic association between minerals were studied in detail. A systematic lead flotation testwork program was carried out to obtain the optimal flotation and separation conditions, and the products obtained were analyzed. The results show that the concentrate contains a wide variety of minerals with complex material composition, and the lead mineral was mainly galena with a relative content of 3.43% and a particle size −37 μm accounting for 94.72%, while the gold minerals were dominated by electrum. The grades of gold, silver, and lead in the balland obtained through the flotation closed-circuit test were 512.10 g/t, 1632.80 g/t, and 40.38%, and the recoveries were 70.65%, 73.86%, and 75.37%, respectively. The gold lost in the flotation tailings was mainly dominated by gold encapsulated in metal sulfide (accounting for 55.67%), and the lead lost was mainly in gangue and metal oxides (accounting for 62.72%).

1. Introduction

Lead is an important industrial raw material due to its low melting point [1], high corrosion resistance, and strong X-ray shielding capacity. It is widely used in various industries, including chemical manufacturing and battery production [2,3]. Gold and silver, as two critical precious metals, not only have broad industrial applications but also serve as vital assets for investment and national reserves [4]. These metals are commonly found in their native states or embedded within the structures of sulfide and silicate minerals, frequently coexisting with minerals such as chalcopyrite, galena, sphalerite, and pyrite [5,6]. In most flotation plants, gold is often enriched alongside lead, copper, zinc, and pyrite minerals [7]. With the depletion of high-grade, easily processed ore resources, increasing attention in the mining sector has shifted toward the efficient utilization of complex and refractory polymetallic ores. Among these, lead-bearing polymetallic concentrates have become a research focus due to their significant economic potential [8,9]. These concentrates typically contain lead, gold, silver, and other valuable metals. The efficient separation and enrichment of these metals are not only essential for maximizing economic returns but also play a key role in improving mining efficiency and resource utilization rates [10,11]. Lead–gold mixed rougher concentrates often exhibit complex compositions. In addition to galena (PbS), they commonly contain pyrite (FeS₂), sphalerite (ZnS), arsenopyrite (As₂S₃), and native gold and silver [12,13]. The intricate interlocking of these minerals, combined with their diverse physical and chemical properties, poses significant challenges to traditional mineral processing techniques [14,15]. For instance, lead minerals are often finely disseminated, while gold and silver may occur as ultrafine particles embedded within sulfide minerals [16,17], rendering them difficult to recover via single beneficiation methods [18]. Therefore, integrated separation strategies are required [12].

Depending on the bulk mineralogy and metal deportment, gravity concentration and flotation are often employed prior to chemical treatment [5]. Among them, flotation is the preferred method due to its effectiveness in concentrating and separating fine-grained minerals [19]. However, to optimize flotation performance, a detailed process mineralogical investigation must be conducted in advance to guide reagent selection and parameter settings. As an interdisciplinary field integrating mineralogy, metallurgy, and materials science, process mineralogy has become a powerful theoretical tool for addressing the beneficiation challenges of complex ores [10,20]. Through systematic analysis of mineral composition, particle size distribution, liberation characteristics, texture, and mineral associations, process mineralogy provides essential data for flowsheet design and optimization [21]. Modern analytical techniques such as mineral liberation analysis (MLA), scanning electron microscopy (SEM), and electron probe microanalysis (EPMA) enable quantitative determination of the occurrence states and interactions of valuable minerals. Such information is critical for guiding the selection and adjustment of beneficiation processes including flotation, gravity separation, roasting, and leaching [22,23].

As one of the most commonly and effectively used techniques for treating sulfide ores and their complex counterparts, flotation has been extensively applied in the recovery of galena, gold, and silver minerals [24,25]. However, flotation efficiency is greatly influenced by factors such as mineral liberation, reagent regimes, and process parameters [26]. Previous studies have demonstrated that a deep understanding of mineral surface chemistry, particle size effects, and interlocking relationships is essential for improving flotation performance [3,27]. Luo et al. [28] enhanced lead mineral recovery from low-grade lead–zinc ores by raising slurry concentration from 27% to 55%, increased lead recovery from 60% to 80%, and boosted lead grade from 27.5% to 29.1%. Li et al. [13] proposed a synergistic MIBC + kerosene decarburization before flotation to reduce metal loss, and achieved subsequent flotation recoveries of lead and zinc at 87.64% and 94.09%, respectively. Yu et al. [9] conducted detailed mineralogical characterization and flotation optimization on low-grade oxidized lead–zinc ore. Through the study, they clarified the mineral symbiosis and determined an optimal process route, achieving a lead grade of 2.83% with 57.56% recovery, and a zinc grade of 28.64% with 83.45% recovery. Li et al. [10] employed a gold flotation-concentrate leaching lead–zinc flotation process for low-grade refractory ores. This yielded a gold concentrate with an Au concentration of 40.23 g/t and 86.25% recovery, achieving a 98.76% leaching rate, along with effective recovery of lead and zinc.

The complex lead–gold rougher concentrate investigated in this study was derived from a pre-concentration flotation operation and features a complex mineralogical composition. It primarily includes galena, pyrite, arsenopyrite, electrum, and minor amounts of chalcopyrite and sphalerite. Preliminary chemical analysis revealed that the concentrate contained 3.74% Pb, 50.50 g/t Au, and 154.80 g/t Ag, indicating high economic value. However, most of the gold was either embedded in or encapsulated by fine-grained sulfide minerals, posing significant challenges to traditional flotation and leaching technologies.

In this study, a detailed process mineralogical analysis of the Pb–Au concentrate was carried out using X-ray fluorescence spectrometry (XRF), mineral liberation analysis (MLA), and electron probe microanalysis (EPMA) and polarized light microscopy. This work characterized the mineral composition, particle size, and mineral associations of lead, gold, and silver, providing theoretical support for selecting appropriate flotation methods. Systematic flotation experiments were conducted to determine optimal separation conditions. Additionally, exploratory cyanide leaching tests were performed on the flotation concentrate, and the products were thoroughly evaluated. By integrating process mineralogy with flotation techniques, this study proposes a theoretical and technical framework for the efficient utilization of complex Pb–Au rougher concentrates. The findings offer valuable guidance for the comprehensive recovery of similar refractory polymetallic mineral resources and hold significant academic and engineering implications.

2. Experiment

2.1. Materials

2.1.1. Sample

The lead–gold rougher concentrate material, weighing approximately 200 kg, was received from a company in Liaoning. The sample was dried, blended, split, and bagged for the testwork.

2.1.2. Reagent

Activated carbon and sodium sulfide were used as detackifiers; zinc sulfate, sodium thiosulfate, sodium sulfide, and sodium sulfite were used as inhibitors; sulfur, SN9, butyl ammonium black, aniline black, ethyl yellow, isopropyl yellow, and CG1325 as trapping agents; and MIBC as a foaming agent. Sodium cyanide was used as the leaching agent for gold, while sodium carbonate and calcium oxide were used as slurry conditioners. Tap water was used throughout, and all chemicals were sourced from McLean Biochemical Technology Co. (Shanghai, China). All reagents were of analytical grade, and pH modifiers were used as necessary for the reactions.

2.2. Methods

2.2.1. Process Mineralogy Study

Lead, gold, and silver in the raw materials were characterized using a Mineral Liberation Analyzer (Thermo Fisher Scientific, Waltham, MA, USA), Polarizing Microscope (Eclipse LV100POL, Nikon, Tokyo, Japan), Electron Probe Microanalyzer (EPMA JXA-8230, UZONGLAB, Shanghai, China), and Scanning Electron Microscope (Sigma 300, Kefu Mechanical and Electrical Equipment Co., Ltd., Hangzhou, China). Various particle size fractions were analyzed for mineralogical characteristics and mineral liberation. The chemical composition of the sample was analyzed using an X-ray Fluorescence Spectrometer (EDX-7000, Xinyichuang Technology Co., Ltd., Shenzhen, China) and an Inductively Coupled Plasma Spectrometer (PerkinElmerAvio550, PerkinElmer, Shanghai, China).

To analyze the particle size distribution and morphology of gold minerals, heavy mineral concentrates were first obtained through manual panning (gravity concentration). The recovered concentrate was then observed under a polarizing microscope. Gold particles were manually picked and transferred onto glass slides for imaging. Their sizes were measured using a calibrated eyepiece micrometer, and morphological types were recorded and classified based on visual characteristics (e.g., flaky, angular, elongated). A sufficient number of particles were examined to ensure representative statistical data. This manual method was adopted due to the low content and fine-grained nature of the gold particles, and was effective in capturing their liberation state and surface features.

2.2.2. Flotation

The flotation test was conducted by preparing a slurry with 28% solid concentration (comprising 500 g of ore and water) in a plastic bucket. The slurry was stirred using a mechanical stirrer at 400 rpm for 5 min. After mixing, the slurry was quantitatively transferred to a 1.5 L air-inflated flotation cell for testing. Care was taken to prevent sample stratification during transfer. The flotation machine was operated at 1900 rpm. Depending on the test requirements, reagents such as collector, inhibitor, depressant, regulator, and frother were added, and pH monitoring was conducted during the flotation tests. Flotation was performed at room temperature, with a mixing time of 8 min. The flotation concentrates, middlings, and tailings were dried, weighed, and sampled for analysis of the target element grades. The corresponding yields and recoveries were then calculated.

2.2.3. Leaching

Leaching tests were conducted in a 0.5 L conical flask in the magnetic stirring water bath with a slurry concentration of 33%. A 100 g concentrate sample and water were added, followed by pretreatment with calcium oxide for a specified duration. Sodium cyanide was then added for leaching over a period of 48 h. The leachate and leaching residue were collected. The residue was washed three times before being sent for analysis. The leaching rate (α) was calculated using the following formula: where m₀ and m_t are the ore masses before and after leaching, and w₀ and w_t represent the corresponding metal grades.

$$\alpha =1-{m_{t}w}_t/{m_{0}w}_0$$

(1)

Comparative tests analyzed the fluctuation range of recovery and grade for each factor in the flotation condition tests, assessing the impact of these factors on flotation indices. The fluctuation range was calculated as follows:

$$\gamma ={\gamma }_{max}-{\gamma }_{min}$$

(2)

$$\beta ={\beta }_{max}-{\beta }_{min}$$

(3)

where $\gamma$ represents the recovery rate, fluctuation $\beta$ represents the grade fluctuation, ${\gamma }_{max}$ represents the maximum recovery rate of each factor, and ${\gamma }_{min}$ represents the minimum recovery rate of each factor. ${\beta }_{max}$ is the maximum grading rate of each factor and ${\beta }_{min}$ is the minimum grading rate of each factor [9].

2.3. Process Flow Chart

The process flow chart is presented in Figure 1. The complex and refractory lead–gold rougher concentrate underwent rougher flotation to produce a rougher concentrate and rougher flotation tailings. The rougher concentrate was further cleaned to obtain lead concentrate and cleaner tailings. The balland then underwent preliminary cyanide leaching to produce a gold and pregnant solution enriched in silver and leaching residue.

3. Results and Discussion

3.1. Process Mineralogy Analysis

3.1.1. Chemical Composition

The spectral analysis, multi-element analysis, and lead phase analysis results of the lead–gold rougher concentrate are presented in

Table 1. Quantitative analysis of minerals.

Elements	Au (g/t *)	Ag (g/t)	S (%)	Fe (%)	As (%)	SiO2 (%)	Pb (%)	Zn (%)
Concentration	50.50	154.80	35.89	34.19	10.61	11.06	3.74	1.59
Elements	Al2O3 (%)	C (%)	CaO (%)	MgO (%)	Cu (%)	Sb (%)	Cr (%)	Cd (%)
Concentration	1.02	0.48	0.47	0.31	0.11	0.054	0.043	0.035

* “g/t” (grams per ton) is equivalent to parts per million (ppm) for solid samples.

and

Table 2. Phase analysis of lead minerals.

Phase	Pb/Sulfide	Pb/Oxide	Pb/Total
Concentration (%)	3.36	0.38	3.74
Relative content (%)	89.84	10.16	100.00

, respectively. Chemical composition analysis (Table 1) shows that the concentrations of Au and Ag in the samples were 50.5 g/t and 154.8 g/t, respectively. The concentrations of Pb, Zn, As, S, and SiO₂ were 3.74%, 1.59%, 10.61%, 35.89%, and 11.06%, respectively, indicating that the sample comprised sulfide, arsenate, and silicate minerals. The main recoverable metals were gold, silver, lead, iron, and zinc. This study primarily focuses on the recovery of lead, gold, and silver minerals. Table 2 shows that lead minerals were present as sulfides (89.84%) and oxides (10.16%). These results indicate that the sample was a low-grade lead–zinc–gold–silver mixed sulfide concentrate with significant amounts of iron, silicon, and arsenic.

3.1.2. Composition and Mineralogical Characteristics

Composition and Content

As shown in

Table 3. Mineral composition and content.

Mineral	Content (%)	Mineral	Content (%)
Pyrite	54.09	Quartz	10.67
Arsenopyrite	21.21	Feldspar	2.12
Galena, Boulangerite	4.03	Pyroxenes	1.57
Sphalerite	2.39	Calcite, Carbonates	1.38
Chalcopyrite, Chalcocite, covellite	0.41	Micas	0.51
Tetrahedrite, Freibergite	0.01	Wollastonite, Kaolinite, graphite	0.24
Lead oxide	0.43	—	—
Limonite, Hematite, Magnetite	0.94	—	—
Total	83.51	Total	16.49

, the composition of the rougher concentrate revealed 82.14% metal sulfides, primarily pyrite, and arsenopyrite, with minor galena, sphalerite, chalcopyrite, chalcopyrite, antimony lead sulfide, cuprite, azurite, and zoisite. Metal oxides constituted 1.37%, mainly lead oxides, limonite, hematite, and magnetite. Gangue minerals were predominantly quartz, with lesser amounts of feldspar, carbonate, mica, and small quantities of other gangue minerals. Gangue minerals were composed mainly of quartz, followed by feldspar, pyroxene, carbonate, and mica, with other minerals present in minor amounts.

Particle Size Distribution of Major Metal Sulfides

Pyrite and arsenopyrite were predominantly concentrated in medium-size fractions (−0.074~+0.037 mm), accounting for 64.32% and 64.95% respectively, with 25%–28% in coarse fractions (>0.074 mm). In contrast, chalcopyrite and galena exhibited significant fine-grained characteristics: 86.70% of chalcopyrite and 94.72% of galena occurred in fine fractions (<0.037 mm) (

Table 4. Particle size distribution of metallic minerals (%).

	Size (mm)	+0.100	−0.100~ +0.074	−0.074~ +0.053	−0.053~ +0.037	−0.037~ +0.010	−0.010	Total
Minerals
Pyrite		10.24	18.47	25.74	25.47	14.11	5.97	100.00
Arsenopyrite		7.04	20.53	23.07	25.29	16.59	7.48	100.00
Chalcopyrite		—	—	1.78	11.52	72.62	14.08	100.00
Galena		—	—	—	5.28	74.01	20.71	100.00

).

Mineral Locking Characteristics

As illustrated in Figure 2a–c, pyrite was the predominant metal sulfide in the sample, accounting for 51.16% of the total mineral content. Its particle size predominantly ranged from 0.074 mm to 0.037 mm, appearing mainly as crystalline or semi-crystalline grains. Cubic and pentagonal dodecahedral forms were rare. Pyrite primarily occurs as liberated grains (79.76%), with minor associations with arsenopyrite, chalcopyrite, galena, and sphalerite. In certain cases, chalcopyrite and galena were observed filling fractures in pyrite or being encapsulated by other metal sulfides. Gold minerals were found within or along the edges of pyrite, indicating that pyrite acts as a carrier mineral for gold [29].

As illustrated in Figure 2d,e, arsenopyrite was the primary arsenic-bearing sulfide, comprising 21.21% of the mineral content, with grain sizes ranging from 0.1 mm to 0.037 mm. Its morphology is predominantly rhombic, wedge-shaped, or columnar in semi-autoclastic forms, with granular types being uncommon. Arsenopyrite mainly occurs as individual grains, with some particles locked within gangue or associated with galena and sphalerite. As shown in Figure 2f, arsenopyrite is closely associated with gold minerals, often encapsulating them, and thus acts as a major carrier of gold.

As illustrated in Figure 2g–i, galena accounted for 3.43% of the mineral content, primarily occurring as irregular grains. Some grains occurred in a semi-autonomous state, while others were embedded within or associated with pyrite. Galena particles were mostly smaller than 0.037 mm and were primarily associated with pyrite, chalcopyrite, and sphalerite. Lead oxides were frequently observed along galena grain boundaries (Figure 2j,k), and no gold minerals were detected in association with galena.

According to Figure 2l,m, chalcopyrite and covellite together constituted 0.41% of the total mineral content, whereas tetrahedrite and freibergite accounted for only 0.01%. Chalcopyrite grains were generally <0.037 mm and exhibited a granular morphology. Some chalcopyrite grains were located along galena edges or secondary copper minerals, with minor exsolution dispersed in sphalerite, indicating a close association with other sulfides. Chalcopyrite predominantly occurred as individual grains (Figure 2n). Tetrahedrite and freibergite occurred in trace amounts and were typically associated with or enclosed within galena, chalcopyrite, or sphalerite. Tetrahedrite often contained silver, with concentrations reaching freibergite levels (Figure 2o). The element mapping image in Figure 2p provides a reference for the complex multiphase distribution of ore metals [30].

Using MLA and optical microscopy, the locked association of pyrite, arsenopyrite, and galena was examined and summarized in

Table 5. Association characteristics of pyrite, arsenopyrite, and galena (%).

	Locked	Liberated	Sulfide-linked *	Sulfide-Coated *	Associated with Chalcopyrite	Locked in Chalcopyrite	Total
Minerals
Pyrite		79.76	10.72	4.38	3.06	2.08	100.00
Arsenopyrite		74.20	13.02	6.05	4.75	1.98	100.00
Galena		64.02	16.83	8.36	8.21	2.58	100.00

* “Sulfide-linked” indicates that the mineral is associated with sulfide minerals, but not entirely surrounded or wrapped by them. “Sulfide-coated” means that a mineral has a layer or coating of sulfide minerals on its surface.

[21]. Pyrite, arsenopyrite, and galena mainly existed as liberations, comprising 60%–80% of total minerals, followed by sulfides that were associated or encapsulated within other sulfides. Pyrite, arsenopyrite, and chalcopyrite were associated with or encapsulated by chalcopyrite, accounting for less than 5%. Galena was encapsulated by chalcopyrite (2.58%), and was associated with chalcopyrite at 8.21%.

3.1.3. Process Mineralogy Characteristics of Gold Minerals

Analysis of Minerals and Fineness

Electrum (fineness 634.9‰) constituted 76.18% of gold minerals by abundance, whereas native gold (fineness 938.9‰) contributed 48.37% of the total gold content (

Table 6. Minerals and fineness of gold minerals.

Gold Minerals	Native Gold	Electrum	Total (%)
Relative content (%)	23.82	76.18	100.00
Fineness (‰)	938.9	634.9	—
Gold content (%)	48.37	22.37	70.74
Distribution rate (%)	31.62	68.38	100.00

). Although electrum showed higher distribution proportion in flotation products (68.38% vs. 31.62%), coarse-grained native gold may have been incompletely liberated [31].

Size Distribution of Gold Minerals

As shown in Figure 3a,b and

Table 7. Size distribution of gold minerals.

Size (mm)	+0.074	−0.074~ +0.053	−0.053~ +0.037	−0.037~ +0.01	−0.010	Total
Content (%)	0.36	0.81	2.13	49.34	47.36	100.00

, a dominant 96.70% of gold minerals occurred in fine fractions (<0.037 mm), with ultra-fine particles (<0.010 mm) accounting for 47.36% (Table 7). Coarse gold (>0.074 mm) constituted only 0.36%, indicating that gold primarily hosts in fine sulfides or as free particles. The largest gold particle observed in this test measured 0.21 × 0.32 × 0.10 mm, seen as a plate flake with an unclean surface (Figure 3c).

The Morphology of Gold Minerals

The morphology of gold minerals, shown in

Table 8. The morphology of gold minerals.

Morphology	Slab-Flaky	Elongated Angular Granular	Angular Granular	Shape Angular Granular	Total
Content (%)	47.38	21.67	19.82	11.13	100.00

, was primarily plate-like flakes, elongated angular grains, angular grains, and sharp angular grains, with other morphologies occurring less frequently.

Association of Gold Minerals

The locked association of gold, shown in

Table 9. Locked association of gold minerals.

Locked Association	Gold	Locked			Enclosed in Metallic Sulfides	Enclosed in Gangue	Total
		Arsenopyrite	Pyrite	Gangue
Relative content (%)	44.38	12.52	4.95	1.15	35.69	1.31	100.00

and Figure 4a–c, revealed through selective dissolution tests that 44.38% of gold existed as monolithic gold. Metal sulfide-encrusted gold, primarily encrusted by arsenopyrite and pyrite, accounted for 35.69%. Locked gold accounted for 18.62%, mostly with arsenopyrite (12.52%), with minor amounts of other congenial morphologies. The gold enclosed in gangue comprised 1.31%.

3.1.4. Process Mineralogy Characteristics of Silver Minerals

Content and Distribution

The silver content in Ag-bearing minerals such as tetrahedrite, freibergite, and gold minerals was analyzed by SEM and EPMA compositional analysis, with detailed results shown in

Table 10. Content and distribution of silver minerals.

Silver Minerals	Relative Content (%)	Ag (%)	Metal Volume	Distribution (%)
Tetrahedrite containing silver	0.0013	4.88	63.44	40.98
Freibergite	0.0004	16.79	67.16	43.38
Silver gold	0.0000505	29.27	14.78	9.55
Others	99.9982495	9.42 × 10−6	9.42	6.09
Total	100.00	—	154.80	100.00

. The average silver content was calculated to be 4.88%, 16.79%, and 29.27%, respectively, with the metal distribution of silver in silver-bearing tetrahedrite, freibergite, and gold minerals at 40.98%, 43.38%, and 9.55%, respectively.

Association Characteristics

As shown in

Table 11. Association characteristics of silver minerals.

Locked Association	Leachable	Enclosed in Metallic Sulfides	Enclosed in Metallic Oxide and Gangue	Total
Content (%)	60.03	21.23	18.74	100.00

and Figure 5, the locked association of silver in the rougher concentrate was analyzed. The relative content of leachable silver was 60.03% (“leachable” silver refers to silver particles with exposed surfaces, which are not encapsulated or tightly bound within sulfide or gangue minerals), while metal sulfide-encrusted silver and gangue and metal oxide-encrusted silver accounted for 21.23% and 18.74%, respectively.

3.2. Flotation Tests

3.2.1. Flotation Conditions Tests

Based on the concentrate’s characteristics and mineral composition, multiple roughing and concentrating stages are typically required to obtain high-quality ballands, as well as gold and silver concentrates. Metal recovery and concentrate grade can be significantly enhanced by selecting appropriate trapping, frothing agents, and conditioners, along with optimizing the flotation process and parameters [32,33]. Consequently, a one-factor test was conducted on the rougher flotation conditions. The lead flotation flow for this process is shown in Figure 6, while Figure 7 displays the effects of key test factors on the lead rougher concentrate’s flotation [9].

Five grinding fineness settings, each representing a different percentage of particles at −0.074 mm, were tested as independent variables to assess their impact on lead recovery and grade [31]. As shown in Figure 7a, when 85% of the particles were at −0.074 mm, lead recovery in the rougher concentrate was 80.54%, and the lead grade was 17.41%. Further increases in fineness showed no significant improvement in flotation indices, thus the optimal fineness was set at 85% for −0.074 mm particles. Figure 7b shows that various reagent removal agents had minimal effect on lead recovery, which remained between 80.0% and 81.5%. The type of reagent removal agent influenced the lead grade in the rougher concentrate. Activated carbon provided a higher lead grade of 17.41%, compared to 8.85% with sodium sulfide and 9.04% with water washing, indicating that sulfide and water washing cannot effectively remove the adhered flotation agent. Therefore, activated carbon was selected as the reagent removal agent. Figure 7c shows that as the activated carbon dosage increased, the yield of lead rougher concentrate gradually decreased, while the grade increased. At 2000 g/t, Pb recovery was 80.54%, and the grade reached 17.45%. Further increases in dosage provided minimal improvement, confirming 2000 g/t as the optimal dosage [34].

From the dosage tests for sodium carbonate and calcium oxide (Figure 7d,e), it was observed that Pb recovery and grade in the lead rougher concentrate gradually decreased as the calcium oxide dosage increased. The optimal calcium oxide dosage was set at 200 g/t, yielding a Pb recovery of 80.15% and a grade of 17.98%. For sodium carbonate, the optimal dosage was 1000 g/t. Test results in Figure 7f show that sodium sulfite increased the lead grade but significantly reduced recovery. Other inhibitors improved the lead grade but decreased recovery compared to no inhibitor use. Additional agents were avoided to simplify the process and reduce water use and costs; thus, no inhibitors were added to the flotation process. As shown in Figure 7g, tests on different trapping agents revealed that butylammonium black at 60 g/t yielded a Pb recovery of 82.05% and a grade of 18.82%, while a combination (80 + 40) g/t produced a recovery of 80.15% and a grade of 17.98%. Butylammonium black at 60 g/t outperformed the combined agents, leading to its selection as the sole trapping agent [35]. As shown in Figure 7h, increasing the dosage of butylammonium black improved lead recovery. At 60 g/t, lead recovery and grade reached their peak; further increases to 70 g/t had minimal effect. Thus, the optimal dosage for butylammonium black was set at 60 g/t.

Figure 8 shows the fluctuation in lead recovery and grade with various factors, including grinding fineness, reagent removal agent type, activated carbon dosage, adjuster type and dosage, inhibitor type, collector type, and butylammonium black dosage. The lead recovery rate was primarily affected by grinding fineness, adjuster dosage, inhibitor type, and collector type, while the lead grade was mainly influenced by grinding fineness, reagent removal agent type, and collector type. The high impact of reagent removal agent type, adjuster, inhibitor, and collector types on lead recovery and grade underscores the importance of selecting appropriate flotation chemicals. The trapping agent type had the greatest influence on flotation indices, while its dosage had the least impact.

3.2.2. Comprehensive Conditioning Test

Based on the flotation condition results, a comprehensive test with “one roughing, two-stage scavenging, and three rounds of cleaning” was conducted, as shown in Figure 9. The results from the comprehensive test (

Table 12. The result of the comprehensive test.

Minerals	Yield/%	Pb Grade /%	Pb Recovery/%
Balland	3.52	55.51	52.18
Cleaner tailings 3	1.53	20.93	8.55
Cleaner tailings 2	3.25	9.85	8.55
Cleaner tailings 1	6.87	6.52	11.96
Middlings 1	8.75	2.71	6.33
Middlings 2	3.36	2.15	1.93
Tailing	72.72	0.54	10.49
Rougher concentrate	100.00	3.74	100.00

) show that after three rounds of cleaning, the balland grade reached 55.14%, with a recovery rate of 52.36%. After two-stage scavenging, the lead grade in the tailings was 0.54%, and the recovery rate in the tailings was 10.49% [9].

3.2.3. Flotation Flowchart Development

The closed-circuit flow and conditions are shown in Figure 10, with the recoveries and grades of gold, silver, and lead in the balland and tailings presented in

Table 13. Flotation flowchart results.

Minerals	Yield (%)	Grade			Recovery (%)
		Au (g/t)	Ag (g/t)	Pb (%)	Au	Ag	Pb
Balland	6.98	512.10	1632.80	40.38	70.65	73.86	75.37
Tailing	93.02	15.96	43.37	0.99	29.35	26.14	24.63
Feed	100.00	50.59	154.31	3.74	100.00	100.00	100.00

. The closed-circuit test results indicate that, after three concentrations and two-stage scavenging, the grades of gold, silver, and lead in the balland were 512.10 g/t, 1632.80 g/t, and 40.38%, with recoveries of 70.65%, 73.86%, and 75.37%, respectively. In the tailings, the grades of gold, silver, and lead were 50.59 g/t, 154.31 g/t, and 0.99%, with respective recoveries of 29.35%, 26.14%, and 24.63% [12]. Based on these findings, to reduce grinding costs, closed-circuit tests were performed without regrinding. As shown in

Table 14. Flotation flowchart results (grinding fineness: −0.074 mm 71.02%).

Minerals	Yield (%)	Grade			Recovery (%)
		Au (g/t)	Ag (g/t)	Pb (%)	Au	Ag	Pb
Balland	7.31	468.80	1543.70	37.47	67.77	73.03	73.05
Tailing	92.69	17.58	44.96	1.09	32.23	26.97	26.95
Feed	100.00	50.56	154.52	3.75	100.00	100.00	100.00

, using a three-stage concentration and second scavenger process, the balland grades for gold, silver, and lead were 468.80 g/t, 1543.70 g/t, and 37.47%, with recoveries of 67.77%, 73.03%, and 73.05%, respectively. In the tailings, the grades of gold, silver, and lead were 17.58 g/t, 44.96 g/t, and 1.09%, with recoveries of 32.23%, 26.97%, and 26.95%, respectively. Comparing the results of regrinding and non-regrinding closed-circuit tests, increasing concentrate fineness improved the recovery of gold, silver, and lead, and increased their grades in the balland [36].

3.3. Leaching Test

To investigate the leaching performance of the flotation balland, a cyanide leaching test was conducted, as shown in Figure 11, with results presented in Figure 12. Test results show that under conditions of regrinding fineness at −0.045 mm (95% content), a calcium oxide dosage of 40 kg/t, a sodium cyanide dosage of 40 kg/t, 33% leaching concentration, and 48-h leaching time, the leaching rates of gold and silver were 76.27% and 47.60%, respectively. Further optimization of reagents and test conditions is necessary to improve gold and silver leaching from lead concentrates. Further optimization of chemicals and test conditions is required to enhance gold and silver leaching from the lead concentrate [4,8].

3.4. Product Analysis

3.4.1. Association of Balland

Liberation and locking association of galena, analyzed by MLA, is shown in

Table 15. Liberation and locking association of galena in balland.

Locked Association	Liberated	Sulfide-Linked	Sulfide-Coated	Hyphenated with Gangues and Metal Oxides	Coated by Gangues and Metal Oxides	Total
Content (%)	80.66	7.85	3.84	3.90	3.75	100.00

. Liberated galena accounted for 80.66%, while other morphologies included associations with sulfides, encapsulation by sulfides, and associations with or encapsulation by gangue minerals and metal oxides. The selective dissolution method, combined with microscopic analysis, was used to examine the distribution of gold minerals in the rougher concentrate (85% at −0.074 mm).

Table 16. Liberation and locking association of gold minerals in balland.

Locked Association	Liberated	Arsenopyrite	Sulfide-Coated	Coated by Gangues and Metal Oxides	Total
Content (%)	61.88	10.37	26.52	1.23	100.00

shows that liberated gold and congenial gold made up 61.88% and 10.37%, respectively, while gold encapsulated in metal sulfide, chalcopyrite, and metal oxides accounted for 26.52% and 1.23%.

3.4.2. Flotation Tailings Analysis

Using MLA and microscopic analysis, the loss patterns of lead minerals in the samples were comprehensively assessed (

Table 17. Analysis of lead minerals in flotation tailings.

Morphology	Liberated	Sulfide-Linked	Sulfide-Coated	Hyphenated with Gangues and Metal Oxides	Coated by Gangues and Metal Oxides	Total
Content (%)	6.10	20.29	10.89	51.47	11.25	100.00

). The analysis indicated that lead minerals lost to the tailings were largely associated with gangues and metal oxides, comprising 62.72% of the total, with 51.47% associated with these materials and 11.25% encapsulated by oxidized gangue and metal. Lead minerals associated with sulfides accounted for 20.29% of losses, while liberated lead loss was relatively low at 6.10%. Assay and analysis clarified the locked association of gold minerals in the sample. Detailed results (

Table 18. Analysis of gold minerals in flotation tailings.

Morphology	Liberated	Intergrowth	Enveloped by Metal Sulfides	Coated by Gangues and Metal Oxides	Total
Content (%)	9.24	30.87	55.67	4.22	100.00

) show that gold was primarily encapsulated in metal sulfides (55.67%), followed by congenial gold (30.87%), liberated gold (9.24%), and gold encapsulated by gangue and metal oxides (4.22%) [37].

3.4.3. Multi-Element Analysis of Products

The multi-element analysis results for balland and flotation tailings are presented in

Table 19. Multi-element analysis of balland.

Element	Au (g/t)	Ag (g/t)	Pb	S	Fe	Zn	SiO2	As
Content (%)	512.10	1632.80	40.38	25.69	13.74	7.26	3.40	2.31
Element	C	Cu	Al2O3	MgO	CaO	Sb	Cd	Cr
Content (%)	1.41	0.92	0.62	0.31	0.30	0.26	0.14	0.019

and

Table 20. Multi-element analysis of flotation tailing.

Element	Au (g/t)	Ag (g/t)	Cu	Pb	Zn	Fe	S	Cr
Content (%)	15.69	43.37	0.08	0.99	0.87	36.38	38.15	0.042
Element	As	C	Al2O3	MgO	SiO2	CaO	Sb	Cd
Content (%)	11.46	0.41	0.97	0.27	8.80	0.49	0.033	0.023

. As shown in Table 19, the grades of Pb, Au, and Ag in the balland were 40.38%, 512.10 g/t, and 1632.8 g/t, respectively, with the S content at 25.69%. Other main metals included Zn and Fe, while the harmful impurity As was present at 2.31%. Table 20 shows that the grades of Pb, Au, and Ag in the flotation tailings were 0.99%, 15.69 g/t, and 43.37 g/t, respectively. The S content was 38.15%, with Fe, Si, and As as the other main elements.

4. Conclusions

(1)

Process mineralogical analysis indicated a diverse and complex mineral composition in the rougher concentrate sample. Metal sulfides constituted 82.14% of the mineral content, with pyrite being predominant at 54.09%. Gangue minerals made up 16.49%, mainly quartz at 10.67%. Average grades for gold, silver, and lead were 50.50 g/t, 154.80 g/t, and 3.74%, respectively.

(2)

Lead minerals were primarily galena, constituting 3.43% of the mineral content. Most galena particles were smaller than 0.037 mm, accounting for 94.72%. Lead phase analysis showed a lead oxidation rate of 10.16%.

(3)

Gold minerals were primarily composed of electrum (76.18%). The particle size was mainly fine-grained (49.34%) and micronized gold (47.36%). Gold minerals were primarily liberated gold (44.38%), followed by gold encapsulated in metal sulfides (mainly arsenopyrite and pyrite) at 35.69%. Congenial gold constituted 18.62%, primarily associated with arsenopyrite at 12.52%, with a small amount in other morphologies. Gold encapsulated in gangue accounted for 1.31%.

(4)

Flotation closed-circuit testing showed that with three-stage cleaning and two-stage scavenging, the grades of gold, silver, and lead in the balland were 512.10 g/t, 1632.80 g/t, and 40.38%, with recoveries of 70.65%, 73.86%, and 75.37%, respectively.

(5)

Cyanide leaching tests on balland, with a regrind fineness of 95% at −0.045 mm, a pulp density of 33%, calcium oxide at 40 kg/t, sodium cyanide at 40 kg/t, and a 48-h leaching time, yielded a gold leaching rate of 76.27%.

(6)

An examination of gold and lead loss in flotation tailings revealed that gold minerals were primarily encapsulated in metal sulfides (55.67%), followed by congenial gold (30.87%), liberated gold (9.24%), and a smaller portion encapsulated by gangue and metal oxides (4.22%). Lead minerals were predominantly associated with gangue and metal oxides (62.72%), with 51.47% at attachment and 11.25% as inclusions. Losses of lead associated with sulfides were 20.29%, and liberated lead loss was minimal at 6.10%.