System and method for improving leaching rate of low-grade scheelite

Abstract

The invention discloses a system and method for improving the leaching rate of low-grade scheelite, and relates to the technical field of mineral recycling. S2, pulping; s2, pulping; s4, crushing and activating by a vertical mill; s5, pressure cooking and leaching; s6, carrying out primary solid-liquid separation; s7, enrichment is carried out; and S8, carrying out secondary solid-liquid separation. According to the technical scheme of cooperation of the rotational flow and the vertical mill in the system, heavy and light phases are separated, the efficiency of the vertical mill is improved, the activation energy of ore pulp obtained after ore grinding is increased, the leaching reaction is facilitated, the yield of the whole process is improved, valuable metal enrichment of low-grade raw ore can be achieved, consumption of raw and auxiliary materials is low, multiple times of solid-liquid separation and enrichment are achieved, and the comprehensive utilization rate of the raw ore is improved. Enriched mineral powder and return water are returned to the front section for recycling, so that energy conservation and emission reduction are realized.

Description

Technical Field

[0001] The invention relates to the technical field of mineral recovery, and in particular to a system and method for improving the leaching rate of low-grade scheelite. Background Art

[0002] Tungsten is a rare heavy metal that is difficult to be selected. The commonly used hydrometallurgical method is sodium carbonate leaching. However, due to the addition of more complex dressing agents in the mineral processing process, the subsequent leaching is greatly affected by incomplete defloating during the roasting and defloating process and the nature of the ore itself. The more mature technology can only ensure that the leaching slag is about 0.5%, and some leaching slag still exceeds 1% or even higher; the mineral powder currently used partially agglomerates during the roasting process, and can reach 99% through a 60-mesh sieve, but the matter on a 200-mesh sieve is more than 8%, which also directly affects the leaching effect. The incompletely leached mineral powder is treated as hazardous waste together with the leaching slag, and valuable metals are lost, which has a great impact on the leaching rate. Summary of the invention

[0003] The purpose of the present invention is to overcome the deficiencies of the prior art and provide a system and method for improving the leaching rate of low-grade scheelite, thereby improving the leaching rate of mineral powder.

[0004] The technical solution adopted by the present invention is: A system for improving the leaching rate of low-grade scheelite, comprising a roasting furnace, a pulping tank, a cyclone, a vertical mill, an autoclave, a storage tank, a filter press I, a felt machine, a suspension machine enrichment and a filter press II; The roasting furnace outlet is connected to the pulping tank feeding port through a conveyor, the pulping tank outlet is connected to the cyclone inlet through a pump, the cyclone light phase outlet is connected to the autoclave inlet, the heavy phase outlet is connected to the vertical mill connection port, and the vertical mill outlet is connected to the autoclave inlet through a pump; the autoclave outlet is connected to the storage tank inlet, the storage tank outlet is connected to the filter press I inlet through a pump, the filter press I leached slag outlet is connected to the slurry mixing barrel inlet, the slurry mixing barrel outlet is connected to the felt machine inlet, the felt machine coarse concentrate outlet is connected to the suspension machine inlet, the suspension machine concentrate outlet is connected to the filter press II inlet, the ore powder outlet of the filter press II is connected to the pulping tank, and the water outlet is connected to the filter press I leached slag tank through a pipeline.

[0005] A method for improving the leaching rate of low-grade scheelite, comprising the following specific steps: S1: roasting and defloating; the ore grade is calculated by weight fraction WO 3 The content is 22-30%, the moisture content is 10-15% and enters the roasting furnace for roasting and deflocculation. The roasting temperature is 500-750℃, and the roasting time is 0.5-2 hours. After roasting, COD≤500mg / L, and the moisture content is not more than 0.1%; S2: slurry preparation; the calcined mineral powder is added with water for slurry preparation, and the liquid-solid ratio of water to mineral powder is 3-5:1; sodium carbonate solid with a content of ≥98% and silicon removal agent are added during the slurry preparation process, and the sodium carbonate concentration after slurry preparation is 130-220g / L; S3: cyclone diversion; the slurry prepared in S2 is transported to the cyclone through a pump, the light phase diverted by the cyclone enters the autoclave, and the heavy phase enters the vertical mill; S4: Vertical mill crushing and activation; the heavy phase is crushed and activated in the vertical mill, the grinding time is not less than 30 minutes, and the mineral particle size -60 mesh pass rate is greater than 95%; the slurry after the vertical mill is transported to the autoclave by a pump; S5: pressure cooking and leaching; the light phase separated by the cyclone in S3 and the pulp after the vertical mill in S4 are pressure cooked and leached in the pressure cooker, the pressure cooking temperature is controlled at 185-220°C, and the material is discharged to the storage tank after keeping warm for 0.5-2 hours; S6: Primary solid-liquid separation: the slurry in the storage tank is pumped to the filter press I for solid-liquid separation, and the separated valuable tungsten-containing solution enters the subsequent workshop for processing; S7: Enrichment; add defoamer to the leached residue in S6 to adjust the slurry, the slurry concentration is ≤20% by weight, and after adjustment, it enters the felt machine for primary enrichment, and the coarse concentrate from the felt machine enters the suspension machine for secondary enrichment; the tailings from the two enrichments are both sent to the tailings pond; S8: Secondary solid-liquid separation; the concentrate from the suspension machine enters the filter press II for solid-liquid separation, the ore powder from the filter press II returns to S2, and the return water is added with flocculant and enters S7 for leaching slag slurry adjustment.

[0006] Specifically, the desiliconizing agent in S2 is one or more of magnesium oxide with a content ≥85% by weight, industrial aluminum sulfate, and industrial aluminum hydroxide with a content ≥64.5% by weight, and the amount of desiliconizing agent added is 3-5% of the weight of the mineral powder.

[0007] More specifically, the calcium oxide content in the magnesium oxide is less than 5%.

[0008] Specifically, in S7, the amount of defoaming agent added is 3‱ of the total solution volume.

[0009] Specifically, the amount of flocculant added in S8 is based on the turbidity of the water to supplement polyacrylamide, and the turbidity is required to be ≤70 mg / L.

[0010] Due to the adoption of the above-mentioned technical solution, the present invention has the following advantages: The technical solution of the combination of cyclone and vertical mill in the system of the present invention can separate the heavy phase from the light phase, improve the efficiency of the vertical mill, and the slurry after grinding has increased activation energy, which is beneficial to the leaching reaction, so that the yield of the whole process is improved, and the valuable metal enrichment of low-grade ore can be achieved, the consumption of raw and auxiliary materials is small, multiple solid-liquid separation and enrichment are carried out, and the enriched ore powder and return water are returned to the front stage for reuse, thereby saving energy and reducing emissions. BRIEF DESCRIPTION OF THE DRAWINGS

[0011] Figure 1 It is a schematic diagram of the process of the present invention. DETAILED DESCRIPTION

[0012] The present invention will be further explained below in conjunction with the accompanying drawings and embodiments, which should not be used to limit the protection scope of the present invention. The purpose of disclosing the present invention is to protect all technical improvements within the scope of the present invention.

[0013] Combined with Figure 1 A system for improving the leaching rate of low-grade scheelite is shown, comprising a roasting furnace, a pulping tank, a cyclone, a vertical mill, an autoclave, a storage tank, a filter press I, a felt machine, a suspension machine enrichment and a filter press II; The roasting furnace outlet is connected to the pulping tank feeding port through a conveyor, the pulping tank outlet is connected to the cyclone inlet through a pump, the cyclone light phase outlet is connected to the autoclave inlet, the heavy phase outlet is connected to the vertical mill connection port, and the vertical mill outlet is connected to the autoclave inlet through a pump; the autoclave outlet is connected to the storage tank inlet, the storage tank outlet is connected to the filter press I inlet through a pump, the filter press I leaching slag tank outlet is connected to the pulp mixing barrel inlet, the pulp mixing barrel outlet is connected to the felt machine ore inlet, the felt machine coarse concentrate outlet is connected to the suspension machine ore inlet, the suspension machine concentrate outlet is connected to the filter press II inlet, the ore powder outlet of the filter press II is connected to the pulping tank, and the water outlet is connected to the filter press I leaching slag tank through a pipeline. Example 1

[0014] A method for improving the leaching rate of low-grade scheelite, comprising the following specific steps: S1: roasting and defloating; the ore grade of the upstream beneficiation unit unloaded by the filter press is calculated by weight fraction WO 3 The content is 22%, the moisture content is 10%, the raw ore enters the roasting furnace at a rate of 3t / h for roasting and defloating, the roasting temperature is 500℃, and the roasting time is 2 hours; the grade of the ore powder after roasting is WO 3 It is 22.05%, moisture 0.02%, COD500mg / L.

[0015] S2: Pulping; Pulping was carried out in a 20 cubic meter pulping tank, 3 tons of roasted mineral powder was added, and 15 cubic meters of water was used for pulping; 1970 kg of solid sodium carbonate with a content of ≥98% was added during the pulping process to make the sodium carbonate concentration 130 g / L; magnesium oxide with a content of ≥85% was added at 3% of the weight of the mineral powder, totaling 105.88 kg, and the calcium oxide content in the magnesium oxide was less than 5%; after pulping, the pulp was sampled and passed through a 60-mesh sieve with a pass rate of 72.50%.

[0016] S3: cyclone diversion; the slurry prepared in S2 is transported to the cyclone through a pump, and the light phase diverted by the cyclone enters the autoclave, and the heavy phase enters the vertical mill.

[0017] S4: Vertical mill crushing and activation; the heavy phase is crushed and activated in the vertical mill, the grinding time is 30 minutes, the mineral particle size -60 mesh pass rate is 95.24%, which meets the post-grinding process requirements; if the pass rate of the 60 mesh screen is ≤95%, the grinding time needs to be extended; the qualified slurry after the vertical mill is pumped to the autoclave.

[0018] S5: pressure cooking and leaching; the light phase separated by the cyclone in S3 and the pulp after the S4 vertical mill are pressure cooked and leached in the pressure cooker. The pressure cooking temperature is controlled at 185°C. After keeping warm for 2 hours, the material is discharged to the storage tank.

[0019] S6: Primary solid-liquid separation: The slurry in the storage tank is pumped to the filter press I for solid-liquid separation; the dry weight of the leached residue is calculated to be 2.4 tons after on-site measurement and weighing. 3 The content is 0.53%, the slag recovery rate is 98.077%, and the separated valuable tungsten-containing solution enters the subsequent workshop for processing.

[0020] S7: enrichment; the leached residue in S6 enters the slurry mixing barrel and is added with defoamer for slurry mixing. The defoamer is added at 3‱ of the total solution volume, and the slurry mixing concentration is 20% by weight. After slurry mixing, it enters the blanket machine for primary enrichment. The blanket machine is 100 blankets with a width of 1.0 meters and a length of 9.5 meters. The coarse concentrate from the blanket machine enters two parallel XFJ15-30 suspension machines for secondary enrichment; the tailings from the two enrichments are both sent to the tailings pond; With 200 tons of WO 3 Based on the leaching slag with a content of 0.53%, 2.1 tons of enriched ore powder with a content of 18.20% and a moisture content of 19.85% were obtained after enrichment by a blanket machine and a suspension machine. The direct recovery rate of metal tungsten after verification by the blanket machine and the suspension machine was 28.90%, and the tungsten grade of the tailings was 0.37%. The impact on the overall recovery rate was: the direct recovery rate (slag basis) increased by 0.581%, and the overall recovery rate reached 98.658%.

[0021] S8: Secondary solid-liquid separation; the concentrate enriched by the suspension machine for the second time enters the filter press II for solid-liquid separation, and the ore powder produced by the filter press II returns to S2. The return water is clarified by adding flocculant polyacrylamide. The turbidity is required to be ≤70mg / L before entering S7 for leaching slag slurry adjustment.

[0022] Comparative Example 1 The slurry prepared in step S2 of Example 1 is directly fed into an autoclave and reacted at 185° C. for 2.0 hours. The slurry is then discharged into a filter press for solid-liquid separation to obtain leached residue WO 3 The content is 0.80%; the direct pressure cooking slag recovery rate is 97.098%. Compared with the slag recovery rate of S6 in Example 1, the scheme of Example 1 recovers 6.48 kg more tungsten, and the yield of Example 1 is increased by 0.979% over that of Comparative Example 1.

[0023] Take 200 tons of 0.80% leached slag of Comparative Example 1, and also use step S7 of Example 1 to obtain 4.2 tons of 15.63% enriched ore powder after enrichment by a suspension machine and a felt machine, with a moisture content of 20.25%. The direct recovery rate of metal tungsten after the felt machine and suspension machine is 32.72%; the tungsten grade of the tailings is 0.42%, and the impact on the whole stage yield is: the direct recovery rate (slag basis) is increased by 1.378%, and the whole stage yield is 98.476%. Table 1 is a summary table of the yield data of Example 1 and Comparative Example 1.

[0024] Table 1 category Content of leaching residue in the first solid-liquid separation Direct recovery rate of slag from solid-liquid separation Enrichment increases direct yield of the whole segment Total direct yield Comparative Example 1 0.80% 97.098% 1.378% 98.476% Example 1 0.53% 98.077% 0.581% 98.658% Example 2

[0025] A method for improving the leaching rate of low-grade scheelite, comprising the following specific steps: S1: roasting and defloating; the ore grade of the upstream beneficiation unit unloaded by the filter press is calculated by weight fraction WO 3 The content is 30%, the moisture is 15%, the raw ore enters the roasting furnace at a rate of 3t / h for roasting and defloating, the roasting temperature is 750℃, and the roasting time is 0.5 hours; the grade of the ore powder after roasting is WO 3 The content is 30.09%, moisture 0.02%, COD500mg / L.

[0026] S2: Pulping; Pulping was carried out in a 20 cubic meter pulping tank, 3 tons of roasted mineral powder was added, and 9 cubic meters of water was used for pulping; during the pulping process, 2020.41 kg of solid sodium carbonate with a content of ≥98% was added to make the sodium carbonate concentration of 220 g / L; magnesium oxide with a content of ≥85% was added at 3% of the weight of the mineral powder, totaling 105.88 kg, and the calcium oxide content in the magnesium oxide was less than 5%; aluminum sulfate with a content of ≥64.5% was added at 3% of the weight of the mineral powder, totaling 139.53 kg; after pulping, the pulp was sampled and passed through a 60-mesh sieve with a pass rate of 77.86%.

[0027] S3: cyclone diversion; the slurry prepared in S2 is transported to the cyclone through a pump, and the light phase diverted by the cyclone enters the autoclave, and the heavy phase enters the vertical mill.

[0028] S4: Vertical mill crushing and activation; the heavy phase is crushed and activated in the vertical mill, the grinding time is 30 minutes, and the mineral particle size -60 mesh pass rate is 96.20%; it meets the post-grinding process requirements. If the pass rate of the 60-mesh screen is ≤95%, the grinding time needs to be extended; the slurry after the vertical mill is transported to the autoclave by a pump.

[0029] S5: pressure cooking and leaching; the light phase separated by the cyclone in S3 and the pulp after the vertical mill in S4 are pressure cooked and leached in the pressure cooker. The pressure cooking temperature is controlled at 220°C. After keeping warm for 0.5 hours, the material is discharged to the storage tank.

[0030] S6: Primary solid-liquid separation: The slurry in the storage tank is pumped to the filter press I for solid-liquid separation. After on-site measurement and weighing, the dry weight of the leached residue is calculated to be 2.4 tons. 3 The content is 0.36%, and the slag recovery rate is 99.04%. The separated valuable tungsten-containing solution enters the subsequent workshop for treatment.

[0031] S7: Enrichment; The leached residue in S6 enters the slurry mixing barrel and is added with defoaming agent for slurry mixing. The defoaming agent is added at 3‱ of the total solution volume, and the slurry mixing concentration is 20% by weight. After slurry mixing, it enters the blanket machine for primary enrichment. The blanket machine is 100 blankets with a width of 1.0 meter and a length of 9.5 meters. The coarse concentrate from the blanket machine enters two parallel XFJ15-30 suspension machines for secondary enrichment; the tailings from the two enrichments are both sent to the tailings pond.

[0032] With 200 tons of WO 3 Based on the leaching slag with a content of 0.36%, 1.25 tons of enriched ore powder with a content of 22.08% and a moisture content of 17.95% were obtained after enrichment by a felt machine and a suspension machine. The direct recovery rate of metal tungsten was 31.45%, and the tailings grade was 0.23%. The impact on the overall recovery rate was as follows: the direct recovery rate (slag basis) increased by 0.35%, and the overall recovery rate reached 99.39%.

[0033] S8: Secondary solid-liquid separation; the concentrate from the suspension machine enters filter press II for solid-liquid separation, and the ore powder from filter press II returns to S2. The return water is clarified by adding flocculant polyacrylamide. The turbidity is required to be ≤70mg / L before entering S7 for leaching slag slurry adjustment.

[0034] Comparative Example 2 The slurry prepared in step S2 of Example 2 is directly fed into an autoclave and reacted at 220° C. for 0.5 hours. The discharged material is directly fed into a filter press for solid-liquid separation to obtain leached residue WO 3The content is 0.63%; the direct pressure cooking slag recovery rate is: 98.32%. Compared with the slag recovery rate of S6 in Example 2, the scheme of Example 2 recovers 6.48 kg more tungsten, and the yield of Example 2 is increased by 0.72% over that of Comparative Example 2.

[0035] Comparative Example 2 WO 3 200 tons of leached residue with a content of 0.63% are used, and WO is obtained after enrichment by a suspension machine and a felt machine in the same manner as step S7 of Example 2. 3 The content is 18.20% and the enriched ore powder is 2.7 tons. The moisture content is 18.88%. The direct recovery rate of metal tungsten is 31.64%. The tailings grade is 0.42%. The impact on the overall recovery rate is: the direct recovery rate (slag basis) is increased by 0.56%, and the overall recovery rate reaches 98.88%. Table 2 is a summary table of the recovery data of Example 2 and Comparative Example 2.

[0036] Table 2 category Content of leaching residue in the first solid-liquid separation Direct recovery rate of slag from solid-liquid separation Enrichment increases direct yield of the whole segment Total direct yield Comparative Example 2 0.63% 98.32% 0.56% 98.88% Example 2 0.36% 99.04% 0.35% 99.39%

[0037] It can be concluded from the above two groups of implementation examples that the technical solution of the present invention, which is separated by a cyclone and crushed and activated by a vertical mill before entering a pressure cooker for leaching, and the leached residue is further entered into a felt machine and a suspension machine for enrichment, can recover valuable metals from low-grade ore and greatly improve the direct recovery rate compared to the technical solution of directly entering a pressure cooker for leaching after pulping.

[0038] The parts not described in detail in this invention are prior art.

[0039] The embodiments selected herein for the purpose of disclosing the invention are currently considered to be suitable, but it should be understood that the invention is intended to include all changes and modifications of the embodiments that fall within the scope of the concept and invention.

Claims

1. A system for improving the leaching rate of low-grade scheelite, characterized in that: It includes roasting furnace, pulping tank, cyclone, vertical mill, autoclave, storage tank, filter press I, felt machine, suspension machine enrichment and filter press II; The roasting furnace outlet is connected to the pulping tank feeding port through a conveyor, the pulping tank outlet is connected to the cyclone inlet through a pump, the cyclone light phase outlet is connected to the autoclave inlet, the heavy phase outlet is connected to the vertical mill connection port, and the vertical mill outlet is connected to the autoclave inlet through a pump; the autoclave outlet is connected to the storage tank inlet, the storage tank outlet is connected to the filter press I inlet through a pump, the filter press I leached slag outlet is connected to the slurry mixing barrel inlet, the slurry mixing barrel outlet is connected to the felt machine inlet, the felt machine coarse concentrate outlet is connected to the suspension machine inlet, the suspension machine concentrate outlet is connected to the filter press II inlet, the ore powder outlet of the filter press II is connected to the pulping tank, and the water outlet is connected to the filter press I leached slag tank through a pipeline.

2. A method for improving the leaching rate of low-grade scheelite, characterized in that: The specific steps are: S1: Roasting and defloating; the ore grade is 22-30% WO3 content by weight, and the moisture content is 10-15%. It enters the roasting furnace for roasting and defloating. The roasting temperature is 500-750℃, and the roasting time is 0.5-2 hours. After roasting, COD is ≤500mg / L, and the moisture content is not more than 0.1%; S2: pulping; The calcined mineral powder is added with water for slurry preparation, and the liquid-solid ratio of water to mineral powder is 3-5:1; sodium carbonate solid with a content of ≥98% and silicon removal agent are added during the slurry preparation process, and the sodium carbonate concentration after slurry preparation is 130-220g / L; S3: cyclone diversion; the slurry prepared in S2 is transported to the cyclone through a pump, the light phase diverted by the cyclone enters the autoclave, and the heavy phase enters the vertical mill; S4: Vertical mill crushing and activation; the heavy phase is crushed and activated in the vertical mill, the grinding time is not less than 30 minutes, and the mineral particle size -60 mesh pass rate is greater than 95%; the slurry after the vertical mill is transported to the autoclave by a pump; S5: pressure cooking and leaching; the light phase separated by the cyclone in S3 and the pulp after the vertical mill in S4 are pressure cooked and leached in the pressure cooker, the pressure cooking temperature is controlled at 185-220°C, and the material is discharged to the storage tank after keeping warm for 0.5-2 hours; S6: Primary solid-liquid separation: the slurry in the storage tank is pumped to the filter press I for solid-liquid separation, and the separated valuable tungsten-containing solution enters the subsequent workshop for processing; S7: Enrichment; add defoamer to the leached residue in S6 to adjust the slurry, the slurry concentration is ≤20% by weight, and after adjustment, it enters the felt machine for primary enrichment, and the coarse concentrate from the felt machine enters the suspension machine for secondary enrichment; the tailings from the two enrichments are both sent to the tailings pond; S8: Secondary solid-liquid separation; the concentrate from the suspension machine enters the filter press II for solid-liquid separation, the ore powder from the filter press II returns to S2, and the return water is added with flocculant and enters S7 for leaching slag slurry adjustment.

3. The method for improving the leaching rate of low-grade scheelite according to claim 2, characterized in that: The desiliconizing agent in S2 is one or more of magnesium oxide with a content of ≥85% by weight, industrial aluminum sulfate, and industrial aluminum hydroxide with a content of ≥64.5% by weight, and the amount of desiliconizing agent added is 3-5% of the weight of the mineral powder.

4. The method for improving the leaching rate of low-grade scheelite according to claim 3, characterized in that: The calcium oxide content in the magnesium oxide is less than 5%.

5. The method for improving the leaching rate of low-grade scheelite according to claim 2, characterized in that: In the S7, the amount of defoamer added is 3‱ of the total solution volume.

6. The method for improving the leaching rate of low-grade scheelite according to claim 2, characterized in that: The amount of flocculant added in S8 is based on the turbidity of the water to supplement polyacrylamide, and the turbidity requirement is ≤70mg / L.

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