Deep Processing of Dump Slag from the Copper-Nickel Industry

Abstract

This work proposes an environmentally safe and economically feasible method of waste copper-nickel production slag utilization (Kola Mining-Metallurgical Company, Nornickel, Kola Peninsula, Russia). This process involves the decomposition of slag by diluted solutions of sulfuric acid (7–10 wt.% H₂SO₄) with a transfer in a solution of more than 70% silicon, 77% iron and 78% magnesium, and a concentration of non-ferrous metals in the residue (~70%). Copper ions were used in the leaching stage to prevent the release of hydrogen sulfide into the working atmosphere. Dehydration of the solution, followed by washing of water-soluble sulphates from silica, was carried out to separate silica from the leaching solution. The dehydration temperature effect on the silica structural characteristics was determined. The possibility of recovering non-ferrous metals from solutions after silica extraction by precipitation, in the form of copper cementite, and the sum of nickel and cobalt sulfides, was evaluated. Pigment-grade iron dioxide, magnesium sulphate and aluminium hydroxide were obtained by dehydration of the solution after extraction of base metals, calcination and other operations. Sulfuric acid leaching resulted in the disclosure of sulfide grains encapsulated in ferrosilicate, which is a favorable factor for flotation. The depleted residue can be successfully used in the construction industry.

1. Introduction

The Pechenganickel Mining & Metallurgical Combine (Kola Mining-Metallurgical Company (KMMC), Nornickel, Moscow, Russia) is located in the northwest part of the Kola Peninsula, Russia. The combine processes copper-nickel ore at two industrial sites in the town of Zapolyarny and the settlement of Nickel. Until 2021, it was enriched and metallurgically processed to high-grade matte. Production was associated with sulfur dioxide gases in the atmosphere. Sulfuric gas interacting in the atmosphere with water and air oxygen resulted in the precipitation of already weak sulfuric acid solutions on the ground. It resulted in the leaching of heavy non-ferrous metals from the slag to form solutions in which the content of the nickel and copper significantly exceeded the maximum allowable concentrations. To reduce its environmental impact, the Nornickel Company resolved to abandon the obsolete and environmentally unsound pyrometallurgical technology for roasting copper-nickel concentrate, and the Pechenganickel melting plant ceased operations in 2020 [1]. However, during the combine’s work, a large amount of solid waste (overburdened rocks, tailings, dump slag) has accumulated, for the disposal of which more than 9800 hectares have been allocated. Sulfides of non-ferrous metals and iron in the waste storage process are oxidized with the formation of sulfuric acid; heavy metals are converted to water-soluble salts, resulting in the acidification of natural waters by mining facilities [2,3,4,5,6,7,8,9]. Hypergenic changes in minerals of technogenic products occur much faster than in natural geological conditions. Such a process intensification is provided by the minerals’ surface activation occurring during milling. Thus, despite the closing of the smelter, the problem of the environmental impact of sulfide waste on the environment—“acid mine drainage” (AMD) or “acid rock drainage” (ARD)—remains relevant [10,11,12,13,14,15,16]. In addition to the environmental aspect, there is the problem of the gradual reduction in mineral reserves of rich and easily enriched ores, so there is a need to develop new and cost-effective waste management methods. In this regard, recovering the metals in them, recycling, and using slags as a substitute for natural resources to produce value-added products seem to be favorable options for managing these types of waste [17,18,19,20,21].

The works [18,22,23,24,25] provide extensive reviews on the known methods of utilizing slags from various metallurgical industries. Thus, a perspective direction for slag utilization could be their use in the construction industry [17,26,27,28,29,30]. However, the direct use of slags in this area leads to a total loss of non-ferrous metals, and a series of technological features of this type of resource, which differ from natural ones, are complicated. Slags have low activity, unstable chemical and mineral composition, and contain heavy metals with hazardous properties, posing an environmental risk. All these points limit the use of slags in the construction industry as inert aggregates and fillers [31,32,33,34,35,36,37].

At present, the main approaches to slag processing can be distinguished: pyrometallurgical, hydrometallurgical (bacterial and chemical leaching, flotation) and combined methods [18,22,23,25,32,35,38,39,40,41,42,43,44,45,46,47,48,49]. However, these studies mainly relate to the processing of slags from copper production, and often focus exclusively on non-ferrous metal recovery. This approach is unprofitable because of the low content of non-ferrous metals in the slag, and the need for subsequent utilization or disposal of residues from primary slag processing. Since about 100 million tons of waste slag from copper and nickel production was accumulated at present, and it contains a low concentration of non-ferrous metals, their cost-effective processing requires using affordable and cheap reagents. In addition, the slag processing should be comprehensive, to extract non-ferrous metals and utilize the main macro components of slag: iron and silicon dioxide.

Previously, the possibility of complex processing of nickel production slags (Southern Urals Nickel Plant PJSC, Orsk, Russia) based on hydrochloric acid leaching of the slag was shown [50]. Hydrochloric acid was selected as a leaching agent, due to the low efficiency of sulfuric acid leaching of converter slag and the lack of own production of sulfuric acid. In the KMMC process, a promising reagent for slag processing is the application of substandard sulfuric acid solutions and commercial sulfuric acid. Previously, we have found that, in the etching of slag with dilute solutions of sulfuric acid already at pH = 2, there is a partial dissolution of slag glass; all slag components are destroyed at pH = 0.5, and silicon begins to pass into the solution. As a result of the destruction of silicate particles, sulfides are also exposed, which creates favorable conditions for their subsequent flotation [51].

In this paper, we proposed a fundamentally new integrated approach to dump slag processing in the copper-nickel industry, allowing complete slag utilization, with the separation of major macro components (iron, non-ferrous metals, silicon and magnesium) and their separation in the form of separate products. The results of large-scale laboratory tests on slag sulfuric acid processing in the continuous mode were presented. We describe the processes for obtaining amorphous silica with different structural characteristics. A new method of iron and magnesium separation with the production of ferrous pigment and magnesium sulfate was proposed. The leachate can be used successfully in the construction industry. The initial slag, undissolved residue, and resulting products were characterized using standard physical and chemical analysis methods.

2. Materials and Methods

2.1. Materials

Experiments were performed on a representative sample (20 kg) of dump slag obtained from the Pechenganickel Combine of KMMC (Pechenganickel, Nickel, Russia). The chemical composition of the sample, in terms of oxides, was found to be as follows (wt.%): 39.23 SiO₂, 32.93 FeO, 0.35 CuO, 0.12 CoO, 0.25 NiO, 5.80 Al₂O₃, 2.41 CaO and 11.91 MgO. All other reagents employed in the experiments were of analytical reagent grade. Iron oxide red 130R (Shandong Hongkun Import&Export Co., Ltd., Qingdao, China) was used as commercial specimen.

2.2. Analysis

The mineralogical and chemical composition of the minerals and solid residue were determined based on X-ray diffraction (XRD 6000, Shimadzu, Kyoto, Japan), scanning electron microscopy (SEM LEO-420, ZEISS, Jena, Germany). The surface area, pore diameter and pore volume were determined according to Brunauer–Emmett–Teller (BET) and Barrett–Joyner–Halenda (BJH) methods from N₂ sorption isotherm using a micrometric (TriStar II 3020, Micromeritics, Norcross, GA, USA) surface area and porosity analyzer. Particle size distribution measurements were performed using the SALD-201V particle size analyzer (Shimadzu, Japan). Colorimetric measurements were conducted using portable spectrophotometer (Capsure RM200, X-Rite, Grand Rapids, MI, USA) to obtain three coordinates of the CIELAB color system containing L*, a* and b*, representing redness–greenness, yellowness–blueness and lightness of sample, respectively [52,53,54]. The determination of non-ferrous and noble metals in the samples was carried out using the atomic absorption spectrometer “HGA” 4100 ZL (Perkin Elmer, Shelton, CT, USA), emission spectrometer ICPS-9000 (Shimadzu, Kyoto, Japan) and mass spectrometer ELAN 9000 DRC-e (PerkinElmer, Shelton, CT, USA). The sulfur content was analyzed using a sulfur analyzer (CS-2000, ELTRA, Haan, Germany).

The density ρ was measured with Ostwald–Sprenge type pycnometers with a bulb volume of 10 cm³ and an internal capillary diameter of about 1 mm. The error in the density measurements was estimated to be within ±0.2 g·cm⁻³. The viscosity η of the solution was measured at atmospheric pressure at 20 ± 1 °C using a calibrated Ubbelohde suspended level viscometer (type 1831-1, Shanghai Glass Instruments Factory, Shanghai, China). The kinematic viscosity ν of the solution was calculated by Equation (1):

$$\mathsf{\nu }={\mathrm{k}}_1 · \mathsf{\tau }-{\mathrm{k}}_2 · \mathsf{\tau },$$

(1)

where ν is the low flow time; k₁ and k₂ are the viscometer constants determined by the calibration fluids and the Hagen-Bach correction factor, respectively. The capillary was 0.56 mm in diameter and 40 mm in length, so the correction factor k₂ could be neglected. The viscosity η was then calculated from the density ρ by Equation (2):

$$\mathsf{\eta }=\mathsf{\nu } · \mathsf{\rho },$$

(2)

The values of the viscosity and density of pure water were obtained from the literature [55]. The error of viscosity measurements was ±0.05 mPa·s.

2.3. Sulfuric Acid Leaching Tests

Dry grinding of slag was carried out in grinding cups on a laboratory roller conveyor with the ratio grinding bodies: slag = 4:1 to ensure a material fineness of less than 80 microns. The process of slag leaching was studied by treatment with sulfuric acid solutions with concentrations of 7–12.5% at a solid–liquid ratio (S:L) = 1:(6–10) at a temperature of 20–40 °C. Slag leaching was realized in a glass reactor (XC-2L) equipped with an agitator and a temperature sensor. The stirrer rotation speed was 650 rpm. The pulp was filtered through a Nutsche filter. The undissolved residue was washed with water on the filter in the S:L = (1–5):10. In the batch process, the slag was loaded simultaneously. In the continuous process, the pulp (slag: water) and sulfuric acid were fed with the help of peristaltic pumps. Figure 1 shows the slag processing unit in continuous mode. At the first stage, 7.0 wt.% sulfuric acid preheated to 40 ± 2 °C was loaded into reactor 5. Next, slag was loaded simultaneously at a ratio of S:L = 1:9 into the reactor. After 55 ± 5 min after the start of the slag leaching, the pulp (slag: water = 1:3) was continuously fed into the reactor from tank 1 using peristaltic pump 2, and from tank 3 using pump 4–10.5 wt.% sulfuric acid solution containing 1.0 g·L⁻¹ copper. Simultaneously, the pulp was unloaded using a peristaltic pump 6 to Nutsche filter 7. From it, the filtrate was fed to tank 8. The pulp and acid were loaded at a rate of 0.48 and 0.87 L·h⁻¹, respectively. The reaction mass was unloaded at a rate of 1.35 L·h⁻¹. Since the reaction of slag decomposition with sulfuric acid is exothermic, there was no need for additional heating of the reaction mixture. Before completion of the process, the pumps were stopped, and the reaction mixture was left stirring for 55 min. The undissolved residue was washed with water on the filter in the S:L = (1–5):10. The filtrate was sent to the dehydration stage.

2.4. Recovery of SiO₂

The solution was subjected to dehydration to extract silicon dioxide from the sulfuric acid-leaching solution. Dehydration of the solution was performed in a spray dryer (BXT-8000ST, Shanghai Glomro Industrial Co., Ltd., Shanghai, China) at 150 ± 1 °C and 270 ± 1 °C. That solution was dehydrated in parallel in the air at 45 ± 2 °C. The powders were a mixture of SiO₂ and metal sulfates (iron, nonferrous metals, magnesium and aluminum). The powders were aqueous washed to separate silicon dioxide from metal sulfates by repulsion at 80 ± 2 °C and S:L = 1:3. The SiO₂ samples were dried to a constant weight at 100 ± 2 °C. The sample obtained by dehydration in the air was designated SiO₂-40. Samples obtained by dehydration in a spray dryer at temperatures 150 and 270 °C were SiO₂-150 and SiO₂-270, respectively. Sample SiO₂-40 was additionally purified from impurities (SiO₂-40C).

2.5. Recovery of Non-Ferrous Metals

The copper cementation from solutions after silicon dioxide separation was carried out as follows: in a glass beaker placed in a thermostat, 1 L of the solution iron powder was added in the ratio Fe: Cu = 5:1 [56,57]. The process was carried out under intensive stirring (1000 rpm) for 0.5 h at 70 ± 1 °C. The solution was filtered, and the cementite was dried at 100 ± 2 °C.

Metal sulfides were precipitated according to the method described by [58]. The difference from the above process was that iron sulfide and magnesium carbonate suspensions were used as precipitants and neutralizing agents, respectively. A 1.0 L solution obtained after copper cementation was added into a reactor with an agitator and pH control. The solution was pre-neutralized with magnesium carbonate emulsion (S:L = 1:10) to pH = 3, to reduce the amount of iron sulfide. Then, to precipitate nickel and cobalt sulfides, iron sulfide suspension (S:L = 1:5) was added to the solution under stirring at a rate such that the pH did not exceed 4.0. The suspension was incubated for 1.0 h at 50 ± 2 °C. After completing precipitation, the suspension was acidified to pH = 1.5 to remove an impurity of co-precipitated iron sulfide from the sulfide concentrate. The sulfide precipitate was filtered, washed with a solution acidified to pH = 3.0, and dried.

2.6. Iron and Magnesium Separation

The solution obtained after the separation of nonferrous metals was subjected to dehydration to separate iron and magnesium. This solution was dehydrated in a spray dryer at 200 ± 1 °C. Sulfate powders were calcined in a Nabertherm RT 50-250/11 tube furnace (Nabertherm GmbH, Lilienthal, Germany) at a temperature of 750, 800 and 850 °C for 3 h. The calcined powders were washed of water-soluble compounds at S:L = 1:5 temperature 80 ± 2 °C for 30 min. As a result, Fe₂O₃ samples were obtained. The Fe₂O₃ specimens calcined at 750, 800 and 850 ± 1 °C were designated Fe₂O₃-750, Fe₂O₃-800 and Fe₂O₃-850, respectively.

3. Results and Discussion

3.1. Characterization of Dump Slag

Granulated slag from the Pechenganickel plant consists of dense and tough black granules with a shell-like breakage and a glassy glitter. According to X-ray phase data, magnesia-iron glass (Mg_1.04Fe_0.96SiO₄) is the main mineral component of the presented slag sample (~71%), ferrosilite was found in smaller amounts (FeSiO₃) (~29%), as was quartz (SiO₂); the sample structure was X-ray amorphous (Figure 2).

Mineralogical investigations revealed that the slag matrix consists of olivine, fayalite and glass. Slag glass contains almost no impurities of non-ferrous metals in an isomorphic state; these impurities are concentrated entirely in sulfide alloy inclusions. According to the electron microprobe analysis, the alloy is a pyrrhotite solid solution containing Fe 49–55%, S 34–39%, Ni 4–8%, Cu 2–6% and Co 0.2–0.3%. Inclusions of sulfides in the slag are distributed unevenly and not in all slag grains. Sulfide inclusions are rounded excretions and complex sulfide formations of a curved shape (Figure 3). The sulfide sizes range from 5 to 20 μm. Larger grains up to 50 µm in diameter can also be observed. In the slag matrix, represented by the glass, skeletal needle-like crystals of olivine are observed to form a spinifex structure (Figure 3A,D).

3.2. Leaching Slag Test

Based on the mineralogical composition of the slag, the process of decomposition of its target soluble forms by sulfuric acid can be represented by the following chemical reactions:

$$(\mathrm{F}\mathrm{e},\mathrm{M}\mathrm{g}){\mathrm{S}\mathrm{i}\mathrm{O}}_4+{2\mathrm{H}}_2{\mathrm{S}\mathrm{O}}_4\to \mathrm{F}\mathrm{e}{\mathrm{S}\mathrm{O}}_4+\mathrm{M}\mathrm{g}{\mathrm{S}\mathrm{O}}_4+{\mathrm{H}}_4\mathrm{S}\mathrm{i}{\mathrm{O}}_4,$$

(3)

$$\mathrm{F}\mathrm{e}{\mathrm{S}\mathrm{i}\mathrm{O}}_3+{\mathrm{H}}_2{\mathrm{S}\mathrm{O}}_4\to \mathrm{F}\mathrm{e}{\mathrm{S}\mathrm{O}}_4+{\mathrm{H}}_2\mathrm{S}\mathrm{i}{\mathrm{O}}_3,$$

(4)

$$\mathrm{M}\mathrm{e}\mathrm{S}+{\mathrm{H}}_2{\mathrm{S}\mathrm{O}}_4\to {\mathrm{M}\mathrm{e}\mathrm{S}\mathrm{O}}_4+{\mathrm{H}}_2\mathrm{S}\uparrow$$

(5)

where Me are Fe, Ni and Co, respectively.

To prevent the release of H₂S into the working atmosphere, we proposed a method of slag processing in the presence of copper ions used in a 140–160% excess with the produced hydrogen sulfide [59]. In the presence of copper ions, copper ions interact with the produced hydrogen sulfide to form copper sulfides with a low solubility equilibrium (K_sp) by the following reaction:

$$\mathrm{C}\mathrm{u}{\mathrm{S}\mathrm{O}}_4+{4\mathrm{H}}_2\mathrm{S}\to \mathrm{C}\mathrm{u}\mathrm{S}+4{\mathrm{S}}^0+4{\mathrm{H}}_2\mathrm{O} ({\mathrm{K}}_{\mathrm{sp}\left(\mathrm{CuS}\right)}=2.3\times {10}^{-48}),$$

(6)

At the first stage of work, the periodic study of sulfuric acid leaching behavior of iron, non-ferrous metals, magnesium and silicon depending on the leaching process time, particle size, acid concentration, S:L ratio and temperature was carried out. The results are shown in Figure 4 and

Table 1. Influence of the main parameters on the extraction of the target slag components in the leaching solution.

Parameter	Extraction (%)
	Si	Fe	Mg	Co	Ni
Particle size. Conditions: CH2SO4 = 7%, S:L = 1:10, tinitial = 20 ± 2 °C, τ = 1 h.
<40 μm	69.9	61.4	75.3	66.4	19.5
40–80 μm	64.0	55.1	68.5	59.8	18.7
>80 μm	60.9	55.1	66.4	57.8	16.8
CH2SO4. Conditions: S:L = 1:10, tinitial = 20 ± 2 °C, τ = 1 h, particle size < 80 μm
7.0 wt.%	64.0	55.1	68.5	59.8	18.7
10.0 wt.%	65.5	56.5	70.9	57.6	17.8
12.5 wt.%	73.6	59.8	74.0	44.5	5.9
S:L ratio. Conditions: CH2SO4 = 7 wt.%, tinitial = 20 ± 2 °C, τ = 1 h, particle size < 80 μm
1:15	83.0	88.7	78.8	84.0	25.0
1:10	65.9	57.4	70.3	60.4	19.5
1:9	64.0	55.1	68.5	59.8	18.7
1:6	27.9	28.4	26.0	30.3	9.6
Initial temperature. Conditions: CH2SO4 = 7 wt.%, S:L = 1:9, τ = 1 h, particle size < 80 μm
20 °C	64.0	55.1	68.5	59.8	18.7
40 °C	75.6	75.5	79.9	75.2	20.5
60 °C	70.66	78.8	82.3	79.1	25.7
80 °C	66.60	82.4	85.6	81.7	29.1

.

As can be seen from the data presented in Figure 2, the recovery of iron, magnesium, nickel, cobalt and aluminum in the leaching solution increases with increasing process duration, and the recovery of silicon and calcium decreases. The decrease in the degree of silicon extraction with increasing process time was caused by partial polycondensation of silicic acid. Reducing the percentage of calcium extraction is associated with a decrease in the solubility of calcium sulfate, and with a rising concentration of other metal sulfates in the leaching solution. Increasing the degree of grinding of the initial slag sample increased the degree of extraction of the target components, while finer grinding (less than 40 μm) significantly reduced the filtration rate. Processing the slag with a sulfuric acid concentration of less than 10% showed that the resulting silica-containing solution was in an aggregation-stable state. Increasing the initial concentration of sulfuric acid leads to a significant increase in leachate viscosity and a risk of premature gelation of the leaching solution. The S:L ratio of less than 6 resulted in insufficient decomposition of the slag and very concentrated silicic acid solutions. It can lead to premature solution coagulation and difficulty separating the solutions from the insoluble residue. The S:L ratio of less than 6 resulted in insufficient decomposition of the slag and very concentrated silicic acid solutions. It leads to excessive consumption of acid and complicates their further processing. Loading the slag into sulfuric acid preheated to 40 °C to improve the performance of the process was found. But further increasing the temperature in the system leads to a decrease in the aggregative stability of the silica-containing solution, and requires increased costs for heating the solution without significantly increasing the degree of dissolution of the silica. As a result, the optimal conditions for slag leaching with sulfuric acid are as follows: C_H2SO4 = 7 wt.%, S:L = 1:9, t_initial = 40 ± 2 °C, particle size < 80 μm, τ = 1 h.

Since the slag leaching process in batch mode is less technological on an industrial scale, the tests of slag decomposition with sulfuric acid in the continuous leaching mode were carried out. The data presented in

Table 2. Valuable component distribution in slag leaching with sulfuric acid solution in continuous mode. Conditions: C_H2SO4 = 7 wt.%, S:L = 1:9, t_initial = 40 ± 2 °C, particle size < 80 μm.

Product	Metal Content (g·L−1/wt.%)								Metal Extraction (%)
	Fe	Co	Ni	S	Si	Mg	Ca	Al	Fe	Co	Ni	S	Si	Mg	Ca	Al
Filtrate	17.8	0.07	0.05	–	12.5	5.4	0.5	2.0	60.3	59.9	19.1	–	59.1	65.2	25.2	56.8
Residue	24.6	0.1	0.4	1.9	17.8	6.9	2.36	3.3	39.2	41.6	80.7	100	39.1	38.8	55.3	43.4

were obtained by processing a 5 kg slag sample in continuous mode. Results showed that carrying out the process in continuous operation makes it possible to achieve similar results obtained in batch mode. In this case, the process productivity increases by reducing the time spent on stops for loading and unloading slag.

To increase the degree of extraction of silicon and iron from the slag, we examined the possibility of a two-stage slag treatment. The second leaching stage was also carried out in the reactor in a continuous mode at S:L = 1:5. The results are presented in

Table 3. Valuable component distribution in slag leaching with sulfuric acid solution in continuous mode (stage 2 of slag leaching). Conditions: C_H2SO4 = 7 wt.%, S:L = 1:5, tinitial = 40 ± 2 °C.

Product	Metal Content (g·L−1/wt.%)								Metal Extraction (%)
	Fe	Co	Ni	S	Si	Mg	Ca	Al	Fe	Co	Ni	S	Si	Mg	Ca	Al
Filtrate	24.2	0.08	0.12	–	12.6	6.8	0.72	1.9	43.3	35.0	13.1	–	31.0	42.9	13.4	25.2
Residue	20.9	0.095	0.48	2.8	18.3	5.6	2.84	3.5	57.5	64.3	81.2	100	69.7	54.5	81.4	70.9

.

As a result of a two-stage slag processing with sulfuric acid, the extractions of components were, in wt.%: 72.7 Si, 77.7 Fe, 73.3 Co, 34.5 Ni, 17 Cu, 78.8 Mg, 69.3 Al and 55.0 Ca. The decrease in slag mass after the first leaching stage was 2.5 times, and after the second stage was 1.5 times; the total slag reduction was 3.7 times.

X-ray phase analysis indicates that ferrosilite (2(Mg_0.56Fe_0.44)O·SiO₂)) and calcium sulfate (Ca(SO₄)(H₂O)₂) form the basis of the cakes after first stage sulfuric acid leaching of slag (Figure 5). After the second leaching stage, the residue was magnesium silicate (Mg₂SiO₄) and quartz (SiO₂).

The mineralogical analysis of residues after the first and second stages of sulfuric acid leaching was carried out (Figure 6). The dissolution of the amorphous component of the slag was found to release sulfide inclusions. According to microprobe analysis after the first leaching step, sulfide particles contain, wt.%: 24.0–35.0 Fe, 3.0–10.0 Cu, 8.0–17.0 Ni, 0.6–1.7 Co and 18.0–36.0 S; after the second, wt.%: 25.0–29.0 Fe, 2.0–6.0 Cu, 7.0–22.0 Ni, 0.6–1.7 Co and 17.0–39.0 S.

Thus, two-stage leaching resulted in the disclosure of sulfide particles encapsulated in ferrosilite. This fact is favorable for the use of, for example, flotation to obtain a concentrate of non-ferrous metals [18,60]. The residue after non-ferrous metal extraction can effectively be used in the construction industry [27,28,29].

3.3. SiO₂ Recovery from Sulfuric Acid Slag Leaching Solutions

In sulfuric acid technology of slag processing, filtrate stability is essential. As a result of the polymerization of silicic acid in these solutions over time, structural formation processes occur, which leads to gelation in the system. Without external influences, gel formation in the solution after leaching can occur during the first hours and within the next three to four days. The first stage of polymerization is the formation of a polysilicon acid sol, followed by particle structure formation with the appearance of a gel. One of the most significant rheological parameters characterizing the transition of a sol into a gel is viscosity, the growth of which precedes the formation of a gel. In this regard, the viscosity of the filtrate obtained in Section 3.2 was studied at the first stage of the study of filtrate processing options. The silica content in the solution was 12.5 g·L⁻¹. The filtrates were kept at room temperature (20 ± 2 °C) for 8–9 days until gelation. The results are shown in

Table 4. Dependence of solution viscosity on leaching solution residence time.

Time (Days)	1	2	3	4	5	6	7	8
Viscosity (mPa·c)	1.80	1.93	2.03	2.14	2.40	3.50	5.46	11.35

.

As can be seen from the presented data, the solution viscosity changes over time, and three sections can be distinguished. The first section of a slight increase in viscosity corresponds to the induction period (up to 4–5 days). The second period is characterized by an intense viscosity increase (more than 4–5 days). The third period is characterized by viscosity fluctuations (the measurement error is more than 20%). A viscosity value of 2.4 mPa·c can be considered the beginning of gelation. Thus, it was established that the solutions obtained by sulfuric acid leaching of slag are stored without gelatinization for 4–5 days. Thus, the relatively high stability of solutions allows one to carry out all technological operations without the risk of gelatinization of solutions.

Silicon dioxide obtained from solutions resulting from sulfuric acid leaching of slag was considered. Solution composition was, in g·L⁻¹: 18.9 Fe, 12.5 Si, 0.4 Cu, 0.05 Ni, 0.07 Co, 5.6 Mg, 1.98 Al and 0.5 Ca. Depending on the dehydration temperature, silicon dioxide with different textural characteristics was found to occur. The morphology of silica particles obtained at different dehydration temperatures is shown in Figure 7. As can be seen, the shape and size of the SiO₂ particles change with increasing dehydration temperature. At a dehydration temperature of 40 °C (sample SiO₂-40), particles of 5–10 μm and large irregularly shaped particles larger than 120 μm are formed. As the dehydration temperature increases, the sample becomes more homogeneous, the particles decrease to 1–7 μm, and a small number of large irregularly shaped fused particles are detected. Treatment of the SiO₂-40 sample with hydrochloric acid does not change the shape and size of the particles.

Figure 8 shows the adsorption–desorption isotherms of N₂ for the SiO₂ samples. Nitrogen adsorption isotherms for all specimens belong to type IV, and are characterized by a hysteresis loop in the region (P/P₀) from 0.4 to 1.0. As seen from the data presented in Figure 8, the isotherm for the SiO₂-40C specimen has an initial section with a less sharp increase at P/P₀ < 0.1. This fact indicates that there are practically no micropores in this sample. For the other materials, the sharp increases at P/P₀ < 0.1 and the narrow H4-type hysteresis loop at higher pressures indicate a combination of micro- and mesopores [38,61,62,63,64].

The textural characteristics of SiO₂ specimens are shown in

Table 5. The textural characteristics of SiO₂ specimens.

Specimen	SBET, m2·g−1	VΣ, cm3	Smicro, m2·g−1	Vmicro αs, cm3·g−1	Dpore, nm
SiO2-40	525	0.548	95.35	0.050	4.02
SiO2-40C	440.65	0.595	20.76	0.008	5.40
SiO2-150	545.81	0.744	97.38	0.051	4.97
SiO2-270	676.62	0.739	55.54	0.027	4.27

. As can be seen, with increasing dehydration temperature, the sample-specific surface area grows with overall pore volume rises, and the micropore fraction decreases from 9.1 to 3.6%. The micropore fraction is less than 1.5% when the SiO₂-40 sample is washed with hydrochloric acid.

Studies of the particle size distribution of SiO₂ samples exposed to ultrasonic dispersion showed that the average particle size decreases with an increase in the specific surface area of the specimens; the smallest particle size has the sample SiO₂-270 (Figure 9). According to the results, the SiO₂-40 sample, with a surface area of 525 m²·g⁻¹, has a particle size of less than 2 to 120 μm, with 75% being less than 60 μm, 50% less than 36 μm and 15% less than 8 μm (Figure 9A). When the SiO₂-40 sample is treated with hydrochloric acid, the particle distribution does not change (data for the SiO₂-40C specimen were not presented). For the SiO₂-150, the particle size varies from 1.2 to 82 μm, with 95% being smaller than 53 μm, 75% smaller than 22 μm, and 41.6% smaller than 8.1 μm (Figure 9B). For sample SiO₂-270, the particle size varies from 1.6 to 49.8 μm, with 49% being less than 7 μm (Figure 9C). The decrease in the average particle size is probably due to the faster kinetics of SiO₂ particle formation with an increasing dehydration temperature.

The chemical composition and the main characteristics of SiO₂ samples are presented in

Table 6. The chemical compositions and the main characteristics of SiO₂ samples.

Index	SiO2-40	SiO2-40C	SiO2-150	SiO2-270
SiO2 (wt.%)	73.1	85.12	74.58	87.00
Mass fraction of water (wt.%)	5.56	5.5	2.06	4.2
Loss on ignition (LOI) (%)	4.70	5.0	5.23	8.4
CaO (wt.%)	1.91	≤0.002	0.182	0.064
Fe2O3 (wt.%)	1.22	0.037	0.886	0.100
Al2O3 (wt.%)	0.19	≤0.02	0.170	0.057
TiO2 (wt.%)	0.310	0.140	0.267	0.184
${\mathrm{S}\mathrm{O}}_4^{2-}$ (wt.%)	3.4	0.018	2.362	0.147
SBET (m2·g−1)	525	440	545	598
Bulk density (g·L−1)	76	38	61	29
pH of 4% dispersion	2.3	4.18	2.79	2.72

. The results presented in the table show that the efficiency of washing metals from SiO₂ samples increases with increasing dehydration temperature. SiO₂-40 specimen was processed with solutions of oxalic (${\mathrm{C}}_{{\mathrm{C}}_2{\mathrm{H}}_2{\mathrm{O}}_4}$ = 5–10%) and hydrochloric (C_HCl = 5–28%) acids at S:L = 1:5, τ = 15 min, t = 80 ± 2 °C (data not shown here). Processing with 5–10% hydrochloric acid solutions proved to be the most effective, as a result of which the total impurity content in the sample does not exceed 0.20% (sample SiO₂-40C).

Thus, as a result of the work found, it was shown that in the sulfuric acid treatment of slag, silica with different qualitative characteristics can be obtained. The dependence of the morphology and surface characteristics of the samples on the dehydration temperature was established. Obtained silica can be used as a thickener, anti-caking agent, carrier for catalysts, and also for obtaining sorbents [64].

3.4. Solution Processing after SiO₂ Extraction

Washing of the dehydrated powder and separation of silica resulted in concentrated solutions of iron, nonferrous metals and magnesium sulfate, with the following composition, g·L⁻¹: 34.0 Fe, 0.82 Cu, 0.099 Ni, 0.13 Co, 10.5 Mg, 0.45 Ca and 3.6 Al. Separation of iron, magnesium, aluminum and nonferrous metals from this solution into separate products was considered. Several existing methods for processing such solutions may include cementation, sulfide or hydroxide precipitation, crystallization and extraction. The arguments for preferential use of cementation and metal sulfide are based on the high degree of metal removal at relatively low pH values, the sparingly soluble nature of sulfide precipitates, favorable dewatering characteristics, and the stability of the metal sulfides formed [57,65].

3.4.1. Non-Ferrous Metals Recovery

The traditional option for removing copper from acidic effluent, mine and quarry waters is cementation with metallic iron, which can result in copper concentrate [57]. As a result of cementation, a copper recovery of more than 99% was achieved, and the residual copper content in the solution was 0.5 ppm. According to XRD data, copper cementite contains copper and copper oxide (Figure 10). The blister copper obtained can be used for smelting in production.

Next, according to the method described in Section 2.5, the solution was purified from nickel and cobalt. The resulting sulfide residue was obtained with the following composition, in wt.%: 94.58 S, 4.16 Fe, 0.37 Si, 0.28 Co, 0.13 Ni, 0.06 Cu and 0.05 Ca. The nickel and cobalt recovery were over 96%. The nickel and cobalt remaining content were, in ppm: 0.15 Co, 0.63 Ni. According to the X-ray phase analysis, the residue was present by elemental sulfur, nickel and cobalt sulfides, and magnetite (Figure 11). This residue can be successfully combined with the flotation concentrate from Section 3.2.

3.4.2. Production of Iron Oxide Powders and Magnesium Sulfate

Iron and magnesium can be separated by hydrolytic precipitation of iron, recrystallization or by electroflotation methods [66,67]. Generally, these methods are based on the pre-oxidation of iron(II) to iron(III). However, these methods are not devoid of disadvantages, since hydrolytic precipitation produces contaminated, hard-to-filter high-watered residues. Crystallization often does not achieve 100% yield of the target component. The need to dilute concentrated solutions is a significant disadvantage when using electroflotation. For separating iron and magnesium from solutions after nonferrous metal extraction, we proposed a principally new approach based on temperature differences in the thermal decomposition of sulfates that are part of the obtained solutions. It is well known that an iron (II) and aluminum sulfate thermal decomposition reaction with the formation of their oxides occurs in the temperature range of 580–750 °C, and the decomposition of magnesium sulfate to form magnesium oxide occurs at temperatures above 1100 °C. The solution obtained after extraction of the nonferrous metals was dehydrated in a spray dryer at 200 °C. The result was a powder containing iron, aluminum, magnesium and calcium sulfates. Figure 12 lists the results of the TG-DTA analysis of the dehydrated powder. According to the thermogravimetric analysis, the thermogram of the sample includes several stages of mass loss. The first stage of heating from room temperature to 250 °C involved the removal of physically adsorbed water. Accordingly, peaks corresponding to the endothermic effect are observed in the specified temperature range. Weight loss at this stage is 21.59%. In the second stage, the mass loss is 40.34%. The DTA curve of the samples exhibits two peaks at 720 and 740 °C. In this temperature range, iron and aluminum sulfates decompose to their oxides.

Table 7. Solution composition and extraction of target components in aqueous leaching solution of calcined sulfate powders.

Temperature, °C	Fe		Mg		Ca		Al
	g·L−1	%	g·L−1	%	g·L−1	%	g·L−1	%
750	0.10	0.17	25.2	97.8	0.71	92.3	0.025	5.4
800	0.011	0.02	22.3	94.6	0.64	89.6	<10−3	-
850	<10−3	-	20.2	92.1	0.58	89.4	<10−3	-

presents the results of aqueous leaching of sulphate powders calcined at different temperatures. The data presented in Table 6 show that the degree of transformation of metal sulfates into their oxides increases with increasing calcination temperature, and the transition of impurity elements (Mg, Ca and Al) into iron oxide powder increases.

It has been found that, with additional washing of iron oxide powders with 2 M H₂SO₄, it is possible to improve the quality of iron oxide powders (

Table 8. The qualitative indicators of iron oxide powders.

Property	Fe2O3 Powders
	Fe2O3 R130	Fe2O3-750	Fe2O3-800	Fe2O3-850
Fe2O3 (wt.%)	≥96	97.1	96.4	96.4
Al2O3 (wt.%)	-	2.9	3.4	3.4
CaO (wt.%)	-	0.03	0.02	0.01
Dispersity (μm)	0.6–25.8	0.29–3.4	0.29–5.24	0.29–5.24
pH of aqueous suspension		4.0	4.1	4.6
Water soluble matter (wt.%)		0.4	0.2	0.1

). Aluminum (up to 3.4 wt.% Al₂O₃) remains the main impurity entering the samples. After calcium removal by recrystallization [68] of the magnesium sulfate solution, with subsequent evaporation of the solution, magnesium sulfate containing 0.005 wt.% CaO was obtained (Table 8).

From particle size distribution analysis of the Fe₂O₃-750 specimen, shown in Figure 13, it was found that 7.2% of the particles have an average volume equivalent spherical diameter of less than 3.28 μm, 24.8% have 20.3 μm, and 98% have less than 3.4 μm. Sample particle sizes of Fe₂O₃-800 and Fe₂O₃-850 vary from 0.289 to 5.24 μm, and with increasing calcination temperature, the fraction of particles larger than 1.065 μm rises negligibly. So, in the interval 1.065–5.24 μm, the particle fraction for the sample Fe₂O₃-800 is 73.45%, for the sample Fe₂O₃-850 is 79.5%, also in the sample Fe₂O₃-850, and the fraction of particles with sizes less than 1.065 μm is reduced by 5%. Thus, as the calcination temperature increases, there is a slight decrease in dispersity associated with particle sintering. It should be noted that the obtained samples are characterized by a lower interval of particle size distribution, as compared to the standard specimen (Fe₂O₃ R130), where the particle distribution varies in the range of 0.596–25.791 μm.

The colorimetric parameters of the Fe₂O₃ samples, obtained at different calcination temperatures, and a commercial specimen of iron oxide pigment R130, were measured (Figure 14,

Table 9. Colorimetric analysis.

Sample	Colorimetric Analysis
	L*	a*	b*
Fe2O3-750	64.59	−16.83	49.12
Fe2O3-800	26.58	14.15	11.97
Fe2O3-850	29.88	17.84	11.22
Fe2O3 R130	29.31	18.86	12.97

). As can be seen from the data presented in Table 8, the samples become reddish as the calcination temperature increases. Thus, if there is a significant contribution of green and yellow color at the temperature of calcination at 750 °C, the reddish coloration increases with increasing temperature of calcination. The sample calcined at 850 °C shows the best tonal results compared to the Fe₂O₃-750 and Fe₂O₃-800 samples, with a high red color contribution and a lower contribution of yellow. The higher blue color contribution in the sample Fe₂O₃-850 compared with the sample Fe₂O₃-R130 is probably due to impurity in the form of aluminum oxide and a smaller particle size. The color difference was 2.1%.

The resulting iron oxide powders can be successfully used with a high degree of probability as, for example, color and anticorrosive pigments [69,70,71,72], since the powders are characterized by homogeneous particle size in a relatively narrow range, the pH of the water extract is in the interval of 4.0–4.6, and the content of the water-soluble substance is less than 1.0 wt.% (Table 7).

3.5. Scheme of Sulfuric Acid Processing of Dump Slag

According to the obtained results, a slag utilization scheme was proposed, which includes slag leaching with a sulfuric acid solution, dehydration of leaching solutions, and washing of dehydrated powder to obtain amorphous silica and a solution containing iron, magnesium, aluminum and non-ferrous metal sulfates (Figure 15). The residue is a flotation to produce a concentrate of non-ferrous metals. The residual can be used in the construction industry. Cementation separates copper, nickel and cobalt from the solution obtained, and precipitates as sulfides. The solution obtained after the separation of non-ferrous metals is subjected to dehydration, followed by washing from water-soluble impurities and received iron oxide powder and the solution containing magnesium and calcium sulfates. Then, the filtrate is recrystallized to separate calcium sulfate. After the crystallization of the solution purified of calcium ions, magnesium sulfate is obtained.

4. Conclusions

(1)

A new method of utilizing waste slag from copper-nickel production was developed. The method includes two-stage slag decomposition with dilute solutions of sulfuric acid (7–10 wt.%) with more than 70% silicon, 77% iron and 78% magnesium in the solution, and a concentration of base metals in the residue, separation of silicon dioxide from leaching solutions by solution dehydration, separation of iron oxide powder and magnesium sulfate by solution dehydration, and subsequent calcination of dehydrated powder;

(2)

The effect of leaching solution dehydration temperature on the morphology and surface characteristics of silica was shown. Silica particle size decreased from 60 to 7 μm when the dehydration temperature was increased from 40 to 270 °C;

(3)

A method of separating iron and magnesium from solutions was developed. The difference in decomposition temperatures of iron and magnesium sulphates was based on the process. The calcination of dehydrated powder at 750–850 °C, followed by water leaching, allows the obtaining of iron dioxide of pigment quality and magnesium sulphate;

(4)

Due to sulfuric acid leaching in the undecomposed residue, sulfide grains encapsulated in ferrosilicate were exposed, which is a favorable factor for flotation. The depleted residue can be successfully used in the construction industry.

5. Patents

Kasikov, A.G.; Shchelokova, E.A.; Timoshchik, O.A.; Budnikova, N.N. Method for processing metallurgical slag Patent 2765974 RU, 7 February 2022.