Method for enriching valuable elements through collaborative smelting of magnesium slag and desulfurized gold concentrate

Abstract

The invention discloses a method for enriching valuable elements by collaborative smelting of magnesium slag and desulfurized gold concentrate, which comprises the following steps: uniformly mixing desulfurized gold concentrate, magnesium slag and solvent powder, granulating and drying to obtain mixed granules; the mixed granules are added into a reduction furnace, a reducing agent is added for reduction smelting, and slag and precious iron alloy are obtained; the method comprises the following steps: crushing the precious iron alloy, and leaching to obtain leaching residues and a leaching solution 1; acid is added into the leaching residues for gold dissolving, leaching liquid 2 is obtained, then sodium sulfite is added into the leaching liquid 2, and gold powder is obtained; adding hydrochloric acid into the leachate 1 to obtain a silver chloride solid and a leachate 3; and heating and pressurizing the leachate 3, and adjusting the pH value to obtain the hematite-form precipitate. According to the method, calcium oxide, silicon oxide and magnesium oxide in the magnesium slag serve as slag formers for reduction smelting of the desulfurized gold concentrate in the high-temperature process, iron alloy is obtained through synergistic reduction smelting, meanwhile, valuable elements such as gold, silver and copper in raw materials are captured, the gold recovery rate is increased to 98% or above under the cyanidation-free condition, the silver recovery rate is increased to 95%, and the iron recovery rate reaches 92%.

Description

Technical Field

[0001] The present invention belongs to the technical field of solid waste resource utilization and pyrometallurgy, and particularly relates to a method for enriching valuable elements by collaboratively smelting magnesium slag and desulfurized gold concentrate. Background Art

[0002] As the mainstream process in gold smelting, cyanidation carries significant environmental risks: the cyanide tailings it produces are classified as hazardous waste (HW48) because they contain highly toxic cyanide and heavy metals. Long-term storage can easily lead to groundwater pollution and ecological damage. Although research has been conducted on alternative systems such as thiourea and thiosulfate for cyanide-free gold extraction, it still faces technical bottlenecks such as high reagent costs, poor leaching selectivity, and insufficient process stability. Although pyrometallurgical gold capture technology can avoid cyanide pollution, the existing process has two core flaws: 1) It relies on copper / lead concentrate as a capture medium, which leads to the risk of secondary heavy metal pollution (such as lead vapor escape); 2) The precious metal carrier function of iron resources in gold concentrate (usually with a content of more than 30wt.%) is not fully developed, resulting in the waste of iron resources in the gold smelting process.

[0003] Magnesium slag, a byproduct of the Pidgeon process, produces an average of 6.5 to 8.0 tons of magnesium slag for every ton of magnesium metal produced. This slag is rich in oxides such as CaO, SiO2, and MgO. Currently, it is primarily landfilled, resulting in resource waste and ecological damage. Therefore, a method for synergizing the resource utilization of magnesium slag with desulfurized gold concentrate for the enrichment of valuable elements is urgently needed. Summary of the Invention

[0004] The purpose of the present invention is to provide a method for the synergistic smelting of magnesium slag and desulfurized gold concentrate to enrich valuable elements. By constructing a calcium-magnesium-silicon-based smelting system, the efficient capture of valuable metals such as gold, silver, and iron is achieved, while simultaneously reducing the utilization of magnesium slag, forming a new clean metallurgical process for the coordinated disposal of cross-industry solid waste. The calcium, silicon, and magnesium oxides in the magnesium slag act as slag-forming agents for the reduction smelting of desulfurized gold concentrate during the high-temperature process. Ferroalloys are obtained through synergistic reduction smelting, while simultaneously capturing valuable elements such as gold, silver, and copper from the raw materials. This solves the problems of cyanide-free gold extraction from gold concentrate and magnesium slag reduction.

[0005] The present invention provides a method for co-smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements, comprising the following steps:

[0006] S1. The desulfurized gold concentrate, magnesium slag and solvent powder are uniformly mixed, granulated and dried to obtain a mixed pellet;

[0007] S2. The mixed pellets are added to a reduction furnace, a reducing agent is added, and oxygen is introduced to control the temperature at 1300 to 1500 ℃ for reduction smelting to obtain slag and precious iron alloy;

[0008] S3. The precious iron alloy is crushed and leached to obtain a leaching residue and a leachate 1;

[0009] S4. Adding acid to the leaching residue to dissolve gold to obtain leachate 2, and then adding sodium sulfite to the leachate 2 to obtain gold powder;

[0010] S5. Hydrochloric acid was added to the leachate 1 and the reaction was allowed to proceed for a certain period of time until the precipitation was complete to obtain a solid silver chloride and a leachate 3;

[0011] S6. The leachate 3 is heated and pressurized, and the pH value is adjusted to obtain a precipitate in the form of hematite.

[0012] Furthermore, in step S1, the solvent is one or more of dolomite and quicklime; the ratio of the desulfurized gold concentrate, magnesium slag and solvent satisfies the quaternary basicity R = (CaO + MgO) / (SiO2 + Al2O3) = 0.8 ~ 1.5, and the mass fraction of MgO is less than or equal to 20%, and the mass fraction of Al2O3 is less than or equal to 10%.

[0013] Preferably, the particle size of the magnesium slag and the desulfurized gold concentrate is 0.02 to 1 mm; and the particle size of the solvent powder is 1 to 10 mm.

[0014] Preferably, the moisture content of the mixed pellets is 2-10%, and the particle size is 6-30 mm.

[0015] Furthermore, in step S2, the reducing agent is coal powder or natural gas; the amount of the coal powder is 80-200 kg / t, based on the weight of the desulfurized gold concentrate; the amount of the natural gas is 30-70 Nm 3 / t, measured by weight of desulfurized gold concentrate.

[0016] Preferably, the pulverized coal is anthracite or coking coal; and the particle size of the pulverized coal is 0.1 to 1 mm.

[0017] The reduction smelting temperature is 1300-1500°C and the time is 1-3 hours. Preferably, the reduction temperature is 1350-1450°C; during the reduction smelting, the oxygen flow rate is 30-120 Nm 3 per ton of gold concentrate.

[0018] Preferably, during the reduction smelting, pulverized coal is sprayed into the reduction furnace, with the amount of pulverized coal being 100-150 kg per ton of gold concentrate, to maintain a reducing atmosphere in the reduction furnace.

[0019] Furthermore, in step S3, the leaching is performed using nitric acid and hydrogen peroxide for gold separation; the concentration of the nitric acid is 20-40%, the concentration of the hydrogen peroxide is 10-30%, the leaching temperature is 30-60°C, the leaching time is 3-12h, and the leaching liquid-solid ratio is (1-5):1.

[0020] Furthermore, in step S4, the gold dissolving is specifically performed by adding aqua regia to the leaching residue, wherein the liquid-solid ratio of the leaching residue to the aqua regia is (1-2):1, the temperature is 60-100°C, and the time is 1-4h; the aqua regia is a mixed solution of 63% concentrated nitric acid and 33% concentrated hydrochloric acid.

[0021] Sodium sulfite is added to the leachate 2, specifically, the concentration of sodium sulfite after addition is 80-100 g / L, sodium carbonate is added to control the pH of the solution to 8-10, the reaction temperature is 60-85° C., and the reaction time is 1-3 hours.

[0022] Furthermore, in step S5, the concentration of hydrochloric acid is 1-3 mol / L, the temperature is 80-90° C., and the precipitation time is 1-6 h.

[0023] Furthermore, in step S6, the heating temperature is 150-200° C., the pressure is 1800 KPa, the oxygen partial pressure is 250 KPa, the solution pH is 4.0-4.5, and the reaction time is 2-6 h.

[0024] Principle of the present invention:

[0025] In existing technology, after sulfuric acid roasting and desulfurization, gold concentrate is typically leached with sulfuric acid to preferentially extract copper. The acid leaching residue is then cyanided to extract gold and silver. This conventional process is lengthy, and the large amounts of iron, lead, and zinc contained in the gold concentrate are difficult to utilize. Furthermore, the cyanidation process produces a large amount of cyanide tailings, posing significant environmental challenges.

[0026] The present invention adopts a molten reduction method and utilizes the iron affinity of gold and copper elements to directionally migrate them into the alloy phase under high temperature and reducing environment to generate gold and copper alloys; silver, lead and zinc have poor solubility in iron at high temperature, and the saturated vapor pressure of silver, lead and zinc elements is low at the melting temperature (1300-1500°C), so silver, lead and zinc will be directionally transferred to the gas phase, thereby realizing the comprehensive recovery of valuable elements.

[0027] In addition, the present invention utilizes the characteristics of high calcium and high silicon in magnesium slag to construct a CaO-SiO2-MgO-Al2O3 quaternary slag type, which converts the high-melting point calcium aluminum feldspar (Ca2Al2SiO7) slag (melting point 1593°C) in conventional smelting into low-melting point calcium magnesium feldspar (Ca2MgSi2O7) slag (melting point 1450°C). The viscosity is low at the smelting temperature (1300-1500°C), which is conducive to the volatilization of silver, lead and zinc elements. At the same time, the appropriate viscosity is also helpful for the separation of slag and gold in the smelting process.

[0028] Beneficial effects of the present invention:

[0029] (1) The present invention utilizes the natural slag-forming properties of CaO-MgO-SiO2 (mass ratio 4:2:3) in magnesium slag to replace the traditional quartz flux, constructs a slag type control range of CaO / SiO2=1.2-1.5, reduces the slag viscosity to 0.5-1.0 Pa·s (1300°C), and realizes efficient separation of metal phase and slag phase;

[0030] (2) The method of the present invention controls the amount of reducing agent added and the smelting temperature, so that heavy metals such as Au and Ag in gold concentrate and Cu and Ni in magnesium slag are simultaneously captured by the Fe-based melt to form a precious iron alloy, achieving a multi-metal recovery rate of Au ≥ 98%, Ag ≥ 95%, Cu ≥ 85%, and Ni ≥ 80%;

[0031] (3) The gold recovery rate of the method of the present invention is increased to more than 98% under cyanide-free conditions, the silver recovery rate is increased to 95%, and the iron recovery rate reaches 92%, opening up a new direction for the resource utilization of magnesium slag. 0.8-1.5 tons of magnesium slag can be consumed by processing 1 ton of desulfurized gold concentrate. BRIEF DESCRIPTION OF THE DRAWINGS

[0032] Figure 1 It is a process flow chart of the present invention.

[0033] Figure 2 This is the XRD diagram of the desulfurized gold concentrate.

[0034] Figure 3 This is the XRD pattern of the magnesium slag.

[0035] Figure 4 This is the XRD pattern of the slag in Example 2.

[0036] Figure 5 This is the XRD pattern of hematite in Example 2. DETAILED DESCRIPTION

[0037] Unless otherwise stated, the main chemical components of the raw materials used in the following examples and comparative examples are as follows:

[0038] Desulfurized gold concentrate: CaO 12.31wt.%, SiO2 21.73wt.%, MgO 5.54wt.%, Fe2O3 33.17wt.%, Al2O3 4.73wt.%, CuO 1.87wt.%, PbO 1.77wt.%, ZnO 0.26wt.%, Au 58.97g / t, Ag 130.87g / t. XRD of desulfurized gold concentrate is as follows: Figure 2 shown.

[0039] Magnesium slag: CaO 58.54%, SiO2 27.32%, MgO 6.74%, Fe2O3 2.53%, Al2O3 1.58%, NiO 0.36%. XRD of magnesium slag is as follows Figure 3 shown.

[0040] Comparative Example 1: Taking the traditional sulfuric acid leaching-cyanide leaching process as an example, the desulfurized gold concentrate was acid-leached to extract copper. The acid leaching conditions were: sulfuric acid concentration 1wt.%, solid-liquid ratio 1:4, temperature 60°C, and time 2h. The obtained acid leaching residue was subjected to cyanide leaching. The cyanide leaching conditions were: sodium cyanide dosage 1kg / t, and leaching time 24h. The final gold, silver, and copper recovery rates were 97%, 70%, and 80%, respectively.

[0041] Example 1: The present invention provides a method for smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements. The schematic diagram of the process is as follows: Figure 1 As shown, desulfurized gold concentrate and magnesium slag are used as raw materials, and the particle size after fine grinding and screening meets the particle size of magnesium slag and desulfurized gold concentrate of 0.02-1mm, as follows:

[0042] 1) Ore blending conditions: The mass ratio of desulfurized gold concentrate and magnesium slag is 1:0.8, and dolomite (CaMg(CO3)2) is used as a flux for mixing and granulation. The material ratio satisfies R((CaO+MgO) / (SiO2+Al2O3))=0.8, w(MgO)=10%, w(Al2O3)=2.5%, and the moisture content of the material after drying is 2.43%;

[0043] 2) Reduction smelting conditions: The amount of coal powder added is 120kg per ton of gold concentrate, the pellets are fed into the reduction furnace and smelted at 1400℃ for 1h, and the oxygen flow rate is 30Nm 3 For every ton of gold concentrate, lead, zinc and silver are recovered from the dust, and the pig iron alloy obtained by quenching with molten iron is finely ground to 200 mesh;

[0044] 3) Leaching conditions: 30% nitric acid, 8% hydrogen peroxide, a liquid-to-solid ratio of 3:1, and a leaching temperature of 50°C for 6 hours. Filtering and adding aqua regia at 80°C to adjust the liquid-to-solid ratio to 1.5:1, leaching for 4 hours. Sodium sulfite was added to leachate 2 at a concentration of 100 g / L, the solution pH was 2.0, and the reaction temperature was 85°C for 2 hours to obtain high-purity gold powder with a gold grade of 99.98%. Hydrochloric acid was added to leachate 1 until no precipitation occurred, and AgCl solid was obtained by filtration. The remaining leachate 3 was adjusted to pH 4.0, and precipitation was carried out at 180°C under a controlled pressure of 1800 kPa and an oxygen partial pressure of 250 kPa for 3 hours. Hematite was then filtered to obtain the hematite. The overall process yielded a gold recovery of 97.36%, a silver recovery of 93.24%, a copper recovery of 91.14%, a nickel recovery of 83.54%, an iron recovery of 94.17%, a lead recovery of 93.26%, and a zinc recovery of 98.63%.

[0045] Example 2: The present invention provides a method for smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements. The desulfurized gold concentrate and magnesium slag are used as raw materials. After fine grinding and screening, the particle size of the magnesium slag and desulfurized gold concentrate is 0.02-1 mm. The specific method is as follows:

[0046] 1) Ore blending conditions: The mass ratio of desulfurized gold concentrate and magnesium slag is 1:1, and dolomite (CaMg(CO3)2) is used as a flux for mixing and granulation. The material ratio satisfies R((CaO+MgO) / (SiO2+Al2O3))=1.0, w(MgO)=10%, w(Al2O3)=2.5%, and the moisture content of the material after drying is 2.62%;

[0047] 2) Melting reduction conditions: The amount of coal powder added is 120kg per ton of gold concentrate, the pellets are fed into the reduction furnace and smelted at 1400℃ for 1h, and the oxygen flow rate is 30Nm 3 For every ton of gold concentrate, the slag XRD is as follows Figure 4 As shown, lead, zinc and silver are recovered from the smoke, and the pig iron alloy obtained by quenching the molten iron is finely ground to 200 mesh;

[0048] 3) Stepwise leaching conditions: 30% nitric acid concentration, 8% hydrogen peroxide, liquid-solid ratio 3:1, leaching temperature 50°C, leaching for 6h, filtering and adding aqua regia at 100°C to adjust the liquid-solid ratio to 2:1, leaching for 4h, adding sodium sulfite to leachate 2 at a concentration of 100g / L, solution pH = 2.0, reaction temperature 85°C, reaction time 2h to obtain high-purity gold powder with a gold grade of 99.94%, leachate 1 added hydrochloric acid until no precipitation occurs, filtered to obtain AgCl solid, the remaining leachate 3 adjusted to pH 4.0, controlled pressure 1800kPa, oxygen partial pressure 250kPa precipitation time 3h at 180°C, filtered to obtain hematite, hematite XRD as shown Figure 5 As shown, the gold recovery rate is 99.24%, the silver recovery rate is 95.37%, the copper recovery rate is 91.06%, the nickel recovery rate is 90.33%, the iron recovery rate is 94.34%, the lead recovery rate is 93.74%, and the zinc recovery rate is 98.71%.

[0049] Example 3: The present invention provides a method for smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements. Desulfurized gold concentrate and magnesium slag are used as raw materials. After fine grinding and screening, the particle size of the magnesium slag and desulfurized gold concentrate is 0.02-1 mm. The specific method is as follows:

[0050] 1) Ore blending conditions: The mass ratio of desulfurized gold concentrate and magnesium slag is 1:1.5, and dolomite (CaMg(CO3)2) is used as a flux for mixing and granulation. The material ratio satisfies R((CaO+MgO) / (SiO2+Al2O3))=1.5, w(MgO)=10%, w(Al2O3)=2.5%, and the moisture content of the material after drying is 2.87%;

[0051] 2) Melting reduction conditions: The amount of coal powder added is 120kg per ton of gold concentrate, the pellets are fed into the reduction furnace and smelted at 1400℃ for 1h, and the oxygen flow rate is 30Nm 3 For every ton of gold concentrate, lead, zinc and silver are recovered from the dust, and the pig iron alloy obtained by quenching with molten iron is finely ground to 200 mesh;

[0052] 3) Stepwise leaching conditions: 30% nitric acid, 8% hydrogen peroxide, a liquid-to-solid ratio of 2:1, and a leaching temperature of 50°C for 6 h. The mixture was filtered and aqua regia was added at 100°C to adjust the liquid-to-solid ratio to 1.5:1, and the leaching was continued for 4 h. Sodium sulfite was added to leachate 2 at a concentration of 100 g / L, the solution pH was 2.0, and the reaction temperature was 85°C for 2 h to obtain high-purity gold powder with a gold grade of 99.96%. Hydrochloric acid was added to leachate 1 until no precipitation occurred, and AgCl solid was obtained by filtration. The remaining leachate 3 was adjusted to pH 4.0, and precipitation was carried out at 180°C under a pressure of 1800 kPa and an oxygen partial pressure of 250 kPa for 3 h. Hematite was then filtered to obtain the hematite. The gold recovery was 90.26%, the silver recovery was 80.58%, the copper recovery was 82.17%, the nickel recovery was 80.53%, the iron recovery was 90.73%, the lead recovery was 86.17%, and the zinc recovery was 89.128%.

[0053] Example 4: The present invention provides a method for smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements. The desulfurized gold concentrate and magnesium slag are used as raw materials. After fine grinding and screening, the particle size of the magnesium slag and desulfurized gold concentrate is 0.02-1 mm. The specific method is as follows:

[0054] 1) Ore blending conditions: The mass ratio of desulfurized gold concentrate and magnesium slag is 1:1, and dolomite (CaMg(CO3)2) is used as a flux for mixing and granulation. The material ratio satisfies R((CaO+MgO) / (SiO2+Al2O3))=0.8, w(MgO)=10%, w(Al2O3)=2.5%, and the moisture content of the material after drying is 2.62%;

[0055] 2) Melting reduction conditions: The amount of coal powder added is 120kg per ton of gold concentrate, the pellets are fed into the reduction furnace and smelted at 1400℃ for 1h, and the oxygen flow rate is 20Nm 3 For every ton of gold concentrate, lead, zinc and silver are recovered from the dust, and the pig iron alloy obtained by quenching with molten iron is finely ground to 200 mesh;

[0056] 3) Stepwise leaching conditions: 30% nitric acid, 8% hydrogen peroxide, a liquid-to-solid ratio of 2:1, and a leaching temperature of 50°C for 6 h. The mixture was filtered and aqua regia was added at 100°C to adjust the liquid-to-solid ratio to 1.5:1, and the leaching was continued for 4 h. Sodium sulfite was added to leachate 2 at a concentration of 100 g / L, the solution pH was 2.0, and the reaction temperature was 85°C for 2 h to obtain high-purity gold powder with a gold grade of 99.94%. Hydrochloric acid was added to leachate 1 until no precipitation occurred, and AgCl solid was obtained by filtration. The remaining leachate 3 was adjusted to pH 4.0, and precipitation was carried out at 180°C under a pressure of 1800 kPa and an oxygen partial pressure of 250 kPa for 3 h. Hematite was then filtered to obtain the hematite. The gold recovery was 97.29%, the silver recovery was 96.78%, the copper recovery was 96.16%, the nickel recovery was 86.22%, the iron recovery was 94.03%, the lead recovery was 93.88%, and the zinc recovery was 96.56%.

[0057] Example 5: The present invention provides a method for smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements. The desulfurized gold concentrate and magnesium slag are used as raw materials. After fine grinding and screening, the particle size of the magnesium slag and desulfurized gold concentrate is 0.02-1 mm. The specific method is as follows:

[0058] 1) Ore blending conditions: The mass ratio of desulfurized gold concentrate and magnesium slag is 1:1, and dolomite (CaMg(CO3)2) is used as a flux for mixing and granulation. The material ratio satisfies R((CaO+MgO) / (SiO2+Al2O3))=1.0, w(MgO)=10%, w(Al2O3)=2.5%, and the moisture content of the material after drying is 2.62%;

[0059] 2) Melting reduction conditions: The amount of coal powder added is 120kg per ton of gold concentrate, the pellets are fed into the reduction furnace and smelted at 1350℃ for 1h, and the oxygen flow rate is 30Nm 3 For every ton of gold concentrate, lead, zinc and silver are recovered from the dust, and the pig iron alloy obtained by quenching with molten iron is finely ground to 200 mesh;

[0060] 3) Stepwise leaching conditions: 30% nitric acid, 8% hydrogen peroxide, a liquid-to-solid ratio of 3:1, and a leaching temperature of 50°C for 6 h. The mixture was filtered and aqua regia was added at 100°C to adjust the liquid-to-solid ratio to 2:1, and the leaching was continued for 4 h. Sodium sulfite was added to leachate 2 at a concentration of 100 g / L, the solution pH was 2.0, and the reaction temperature was 85°C for 2 h to obtain high-purity gold powder with a gold grade of 99.90%. Hydrochloric acid was added to leachate 1 until no precipitation occurred, and AgCl solid was obtained by filtration. The remaining leachate 2 was adjusted to pH 4.0, and precipitation was carried out at 180°C under a pressure of 1800 kPa and an oxygen partial pressure of 250 kPa for 3 h. Hematite was then filtered to obtain the hematite. The gold recovery was 82.02%, the silver recovery was 79.13%, the copper recovery was 93.03%, the nickel recovery was 74.23%, the iron recovery was 90.27%, the lead recovery was 80.15%, and the zinc recovery was 83.12%.

[0061] Example 6: The present invention provides a method for smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements. The desulfurized gold concentrate and magnesium slag are used as raw materials. After fine grinding and screening, the particle size of the magnesium slag and desulfurized gold concentrate is 0.02-1 mm. The specific method is as follows:

[0062] 1) Ore blending conditions: The mass ratio of desulfurized gold concentrate and magnesium slag is 1:1, and dolomite (CaMg(CO3)2) is used as a flux for mixing and granulation. The mass ratio of each component of the smelting raw material satisfies R((CaO+MgO) / (SiO2+Al2O3))=1.0, w(MgO)=10%, w(Al2O3)=2.5%, and the moisture content of the material after drying is 2.62%;

[0063] 2) Melting reduction conditions: The amount of coal powder added is 120kg per ton of gold concentrate, the pellets are fed into the reduction furnace and smelted at 1450℃ for 1h, and the oxygen flow rate is 30Nm 3 For every ton of gold concentrate, lead, zinc and silver are recovered from the dust, and the pig iron alloy obtained by quenching with molten iron is finely ground to 200 mesh;

[0064] 3) Stepwise leaching conditions: 30% nitric acid, 8% hydrogen peroxide, a liquid-to-solid ratio of 3:1, and a leaching temperature of 50°C for 6 h. The mixture was filtered and aqua regia was added at 100°C to adjust the liquid-to-solid ratio to 2:1, and the leaching was continued for 4 h. Sodium sulfite was added to leachate 2 at a concentration of 100 g / L, the solution pH was 2.0, and the reaction temperature was 85°C for 2 h to obtain high-purity gold powder with a gold grade of 99.96%. Hydrochloric acid was added to leachate 1 until no precipitation occurred, and AgCl solid was obtained by filtration. The remaining leachate 3 was adjusted to pH 4.0, and precipitation was carried out at 180°C under a pressure of 1800 kPa and an oxygen partial pressure of 250 kPa for 3 h. Hematite was then filtered to obtain the hematite. The gold recovery was 98.98%, the silver recovery was 95.33%, the copper recovery was 97.13%, the nickel recovery was 91.33%, the iron recovery was 95.313%, the lead recovery was 98.233%, and the zinc recovery was 99.86%.

[0065] Example 7: The present invention provides a method for smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements. The desulfurized gold concentrate and magnesium slag are used as raw materials. After fine grinding and screening, the particle size of the magnesium slag and desulfurized gold concentrate is 0.02-1 mm. The specific method is as follows:

[0066] 1) Ore blending conditions: The mass ratio of desulfurized gold concentrate and magnesium slag is 1:1, and dolomite (CaMg(CO3)2) is used as a flux for mixing and granulation. The mass ratio of each component of the smelting raw material satisfies R((CaO+MgO) / (SiO2+Al2O3))=1.0, w(MgO)=10%, w(Al2O3)=2.5%, and the moisture content of the material after drying is 2.62%;

[0067] 2) Melting reduction conditions: The amount of coal powder added is 120kg per ton of gold concentrate, the pellets are fed into the reduction furnace and smelted at 1400℃ for 2h, and the oxygen flow rate is 30Nm 3 For every ton of gold concentrate, lead, zinc and silver are recovered from the dust, and the pig iron alloy obtained by quenching with molten iron is finely ground to 200 mesh;

[0068] 3) Stepwise leaching conditions: 30% nitric acid, 8% hydrogen peroxide, a liquid-to-solid ratio of 2:1, and a leaching temperature of 50°C for 6 h. The mixture was filtered and aqua regia was added at 100°C to adjust the liquid-to-solid ratio to 1.5:1, and the leaching was continued for 4 h. Sodium sulfite was added to leachate 2 at a concentration of 100 g / L, the solution pH was 2.0, and the reaction temperature was 85°C for 2 h to obtain high-purity gold powder with a gold grade of 99.95%. Hydrochloric acid was added to leachate 1 until no precipitation occurred, and AgCl solid was obtained by filtration. The remaining leachate 3 was adjusted to pH 4.0, and precipitation was carried out at 180°C under a pressure of 1800 kPa and an oxygen partial pressure of 250 kPa for 3 h. Hematite was then filtered to obtain the hematite. The gold recovery was 96.18%, the silver recovery was 92.62%, the copper recovery was 89.54%, the nickel recovery was 86.56%, the iron recovery was 92.33%, the lead recovery was 93.26%, and the zinc recovery was 95.37%.

[0069] Example 8: The present invention provides a method for smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements. The desulfurized gold concentrate and magnesium slag are used as raw materials. After fine grinding and screening, the particle size of the magnesium slag and desulfurized gold concentrate is 0.02-1 mm. The specific method is as follows:

[0070] 1) Ore blending conditions: The mass ratio of desulfurized gold concentrate and magnesium slag is 1:0.8, and dolomite (CaMg(CO3)2) is used as a flux for mixing and granulation. The mass ratio of each component of the smelting raw material satisfies R((CaO+MgO) / (SiO2+Al2O3))=0.8, w(MgO)=10%, w(Al2O3)=2.5%, and the moisture content of the material after drying is 2.43%;

[0071] 2) Melting reduction conditions: The amount of coal powder added is 80kg per ton of gold concentrate, the pellets are fed into the reduction furnace and smelted at 1400℃ for 1h, and the oxygen flow rate is 30Nm 3 For every ton of gold concentrate, lead, zinc and silver are recovered from the dust, and the pig iron alloy obtained by quenching with molten iron is finely ground to 200 mesh;

[0072] 3) Stepwise leaching conditions: 40% nitric acid concentration, 10% hydrogen peroxide, liquid-solid ratio 2: 1, leaching temperature 50 ° C, leaching for 8 hours, filtering and adding aqua regia at 100 ° C to adjust the liquid-solid ratio to 1.5: 1, leaching for 4 hours, adding sodium sulfite to leachate 2 at a concentration of 100 g / L, solution pH = 2.0, reaction temperature 85 ° C, reaction time 2 hours to obtain high-purity gold powder with a gold grade of 99.96%, leachate 1 is added with hydrochloric acid until no precipitation occurs, filtered to obtain AgCl solid, the remaining leachate 3 is adjusted to pH 4.0, controlled at 180 ° C. The pressure is 1800 kPa, the oxygen partial pressure is 250 kPa, and the precipitation time is 3 hours, and filtered to obtain hematite. The gold recovery rate of the entire process is 96.76%, the silver recovery rate is 92.58%, the copper recovery rate is 89.33%, the nickel recovery rate is 85.34%, the iron recovery rate is 80.61%, the lead recovery rate is 91.16%, and the zinc recovery rate is 97.53%.

Claims

1. A method for synergistically smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements, characterized in that: The following steps are involved: S1. The desulfurized gold concentrate, magnesium slag and solvent powder are uniformly mixed, granulated and dried to obtain a mixed pellet; S2. The mixed pellets are added to a reduction furnace, a reducing agent is added, and oxygen is introduced to control the temperature to 1300 to 1500 ℃ for reduction smelting and reduction smelting to obtain slag and precious iron alloy; S3. The precious iron alloy is crushed and leached to obtain a leaching residue and a leachate 1; S4. Adding acid to the leaching residue to dissolve gold to obtain leachate 2, and then adding sodium sulfite to the leachate 2 to obtain gold powder; S5. Hydrochloric acid was added to the leachate 1 until no precipitation was produced to obtain a solid silver chloride and a leachate 3; S6. The leachate 3 is heated and pressurized, and the pH value is adjusted to obtain a precipitate in the form of hematite.

2. The method for synergistically smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements according to claim 1, characterized in that: In step S1, the solvent is one or more of dolomite and quicklime; the ratio of the desulfurized gold concentrate, magnesium slag and solvent satisfies the quaternary basicity R = (CaO + MgO) / (SiO2 + Al2O3) = 0.8 to 1.5, and the mass fraction of MgO is less than or equal to 20%, and the mass fraction of Al2O3 is less than or equal to 10%.

3. The method for synergistically smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements according to claim 1, characterized in that: The particle size of the magnesium slag and the desulfurized gold concentrate is 0.02-1 mm; the particle size of the solvent powder is 1-10 mm; the moisture content of the mixed granular material is 2-10%, and the particle size is 6-30 mm.

4. The method for synergistically smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements according to claim 1, characterized in that: In step S2, the reducing agent is coal powder or natural gas; the amount of coal powder is 80-200 kg / t, based on the weight of the desulfurized gold concentrate; the amount of natural gas is 30-70 Nm 3 / t, measured by weight of desulfurized gold concentrate.

5. The method for synergistically smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements according to claim 4, characterized in that: The pulverized coal is anthracite or coking coal; the particle size of the pulverized coal is 0.1-1 mm.

6. The method for synergistically smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements according to claim 1, characterized in that: The reduction smelting temperature is 1300-1500°C and the time is 1-3 hours. During the reduction smelting, the oxygen flow rate is 30-120 Nm 3 per ton of gold concentrate.

7. The method for synergistically smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements according to claim 1, characterized in that: During reduction smelting, pulverized coal is sprayed into the reduction furnace at a rate of 100 to 150 kg per ton of gold concentrate to maintain a reducing atmosphere in the reduction furnace.

8. The method for synergistically smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements according to claim 1, characterized in that: In step S3, the leaching is performed by using nitric acid and hydrogen peroxide to separate the gold; the concentration of the nitric acid is 20-40%, the concentration of the hydrogen peroxide is 10-30%, the leaching temperature is 30-60°C, the leaching time is 3-12h, and the leaching liquid-solid ratio is (1-5):

1.

9. The method for synergistically smelting magnesium slag and desulfurized gold concentrate to enrich valuable elements according to claim 1, characterized in that: In step S4, the gold dissolving is specifically performed by adding aqua regia to the leaching residue, wherein the liquid-to-solid ratio of the leaching residue to the aqua regia is (1-2):1, the temperature is 60-100°C, and the time is 1-4h; the aqua regia is a mixed solution of 63% concentrated nitric acid and 33% concentrated hydrochloric acid.

10. The method for enriching valuable elements by synergistically smelting magnesium slag and desulfurized gold concentrate according to claim 1, characterized in that: Sodium sulfite is added to the leachate 2, specifically, the concentration of sodium sulfite after addition is 80-100 g / L, sodium carbonate is added to control the pH of the solution to 8-10, the reaction temperature is 60-85° C., and the reaction time is 1-3 hours.

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