CN108339658B - Process method for recovering sulfur concentrate from potassium-rich slate - Google Patents
Process method for recovering sulfur concentrate from potassium-rich slate Download PDFInfo
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- CN108339658B CN108339658B CN201810105817.3A CN201810105817A CN108339658B CN 108339658 B CN108339658 B CN 108339658B CN 201810105817 A CN201810105817 A CN 201810105817A CN 108339658 B CN108339658 B CN 108339658B
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- NINIDFKCEFEMDL-UHFFFAOYSA-N Sulfur Chemical compound [S] NINIDFKCEFEMDL-UHFFFAOYSA-N 0.000 title claims abstract description 66
- 239000011593 sulfur Substances 0.000 title claims abstract description 66
- 229910052717 sulfur Inorganic materials 0.000 title claims abstract description 66
- 239000012141 concentrate Substances 0.000 title claims abstract description 57
- 238000000034 method Methods 0.000 title claims abstract description 36
- ZLMJMSJWJFRBEC-UHFFFAOYSA-N Potassium Chemical compound [K] ZLMJMSJWJFRBEC-UHFFFAOYSA-N 0.000 title claims abstract description 35
- 239000011591 potassium Substances 0.000 title claims abstract description 35
- 229910052700 potassium Inorganic materials 0.000 title claims abstract description 35
- 239000010454 slate Substances 0.000 title claims abstract description 27
- MUBZPKHOEPUJKR-UHFFFAOYSA-N Oxalic acid Chemical compound OC(=O)C(O)=O MUBZPKHOEPUJKR-UHFFFAOYSA-N 0.000 claims abstract description 63
- 239000003795 chemical substances by application Substances 0.000 claims abstract description 61
- 238000005188 flotation Methods 0.000 claims abstract description 35
- 230000003213 activating effect Effects 0.000 claims abstract description 29
- 239000004088 foaming agent Substances 0.000 claims abstract description 26
- 239000000203 mixture Substances 0.000 claims abstract description 23
- 150000001412 amines Chemical class 0.000 claims abstract description 21
- 229910000365 copper sulfate Inorganic materials 0.000 claims abstract description 21
- ARUVKPQLZAKDPS-UHFFFAOYSA-L copper(II) sulfate Chemical compound [Cu+2].[O-][S+2]([O-])([O-])[O-] ARUVKPQLZAKDPS-UHFFFAOYSA-L 0.000 claims abstract description 21
- 235000006408 oxalic acid Nutrition 0.000 claims abstract description 21
- KOVPITHBHSZRLT-UHFFFAOYSA-N 2-methylpropoxymethanedithioic acid Chemical compound CC(C)COC(S)=S KOVPITHBHSZRLT-UHFFFAOYSA-N 0.000 claims abstract description 20
- 238000007885 magnetic separation Methods 0.000 claims abstract description 17
- 238000000227 grinding Methods 0.000 claims abstract description 9
- LFQSCWFLJHTTHZ-UHFFFAOYSA-N Ethanol Chemical compound CCO LFQSCWFLJHTTHZ-UHFFFAOYSA-N 0.000 claims description 11
- ZOOODBUHSVUZEM-UHFFFAOYSA-N ethoxymethanedithioic acid Chemical compound CCOC(S)=S ZOOODBUHSVUZEM-UHFFFAOYSA-N 0.000 claims description 11
- 239000012991 xanthate Substances 0.000 claims description 11
- KWYUFKZDYYNOTN-UHFFFAOYSA-M Potassium hydroxide Chemical compound [OH-].[K+] KWYUFKZDYYNOTN-UHFFFAOYSA-M 0.000 claims description 9
- 239000010433 feldspar Substances 0.000 claims description 9
- 229940072033 potash Drugs 0.000 claims description 9
- BWHMMNNQKKPAPP-UHFFFAOYSA-L potassium carbonate Substances [K+].[K+].[O-]C([O-])=O BWHMMNNQKKPAPP-UHFFFAOYSA-L 0.000 claims description 9
- 235000015320 potassium carbonate Nutrition 0.000 claims description 9
- 239000006260 foam Substances 0.000 claims description 7
- 238000002156 mixing Methods 0.000 claims description 3
- 239000002245 particle Substances 0.000 claims 1
- 238000000926 separation method Methods 0.000 abstract description 14
- 239000003153 chemical reaction reagent Substances 0.000 abstract description 4
- 239000010665 pine oil Substances 0.000 abstract description 3
- 229910052500 inorganic mineral Inorganic materials 0.000 description 15
- 239000011707 mineral Substances 0.000 description 15
- 235000010755 mineral Nutrition 0.000 description 15
- 238000003756 stirring Methods 0.000 description 11
- CHWRSCGUEQEHOH-UHFFFAOYSA-N potassium oxide Chemical compound [O-2].[K+].[K+] CHWRSCGUEQEHOH-UHFFFAOYSA-N 0.000 description 10
- 229910001950 potassium oxide Inorganic materials 0.000 description 10
- 238000011160 research Methods 0.000 description 9
- DLHONNLASJQAHX-UHFFFAOYSA-N aluminum;potassium;oxygen(2-);silicon(4+) Chemical compound [O-2].[O-2].[O-2].[O-2].[O-2].[O-2].[O-2].[O-2].[Al+3].[Si+4].[Si+4].[Si+4].[K+] DLHONNLASJQAHX-UHFFFAOYSA-N 0.000 description 7
- 238000011084 recovery Methods 0.000 description 7
- 230000000694 effects Effects 0.000 description 6
- 229910052952 pyrrhotite Inorganic materials 0.000 description 6
- 239000002994 raw material Substances 0.000 description 6
- 229910052626 biotite Inorganic materials 0.000 description 5
- 239000002893 slag Substances 0.000 description 5
- UCKMPCXJQFINFW-UHFFFAOYSA-N Sulphide Chemical compound [S-2] UCKMPCXJQFINFW-UHFFFAOYSA-N 0.000 description 3
- 230000002411 adverse Effects 0.000 description 3
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 2
- 230000007547 defect Effects 0.000 description 2
- 238000005516 engineering process Methods 0.000 description 2
- 239000011019 hematite Substances 0.000 description 2
- 229910052595 hematite Inorganic materials 0.000 description 2
- LIKBJVNGSGBSGK-UHFFFAOYSA-N iron(3+);oxygen(2-) Chemical compound [O-2].[O-2].[O-2].[Fe+3].[Fe+3] LIKBJVNGSGBSGK-UHFFFAOYSA-N 0.000 description 2
- 238000002386 leaching Methods 0.000 description 2
- 238000012545 processing Methods 0.000 description 2
- 239000011028 pyrite Substances 0.000 description 2
- 229910052683 pyrite Inorganic materials 0.000 description 2
- NIFIFKQPDTWWGU-UHFFFAOYSA-N pyrite Chemical compound [Fe+2].[S-][S-] NIFIFKQPDTWWGU-UHFFFAOYSA-N 0.000 description 2
- 229910052665 sodalite Inorganic materials 0.000 description 2
- 229910052889 tremolite Inorganic materials 0.000 description 2
- 239000012190 activator Substances 0.000 description 1
- 229910052612 amphibole Inorganic materials 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 238000001514 detection method Methods 0.000 description 1
- 238000010586 diagram Methods 0.000 description 1
- 238000002474 experimental method Methods 0.000 description 1
- 229910052892 hornblende Inorganic materials 0.000 description 1
- 239000012535 impurity Substances 0.000 description 1
- 239000003112 inhibitor Substances 0.000 description 1
- 229910052742 iron Inorganic materials 0.000 description 1
- 230000001788 irregular Effects 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 239000010453 quartz Substances 0.000 description 1
- 238000004064 recycling Methods 0.000 description 1
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N silicon dioxide Inorganic materials O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 description 1
- 230000002311 subsequent effect Effects 0.000 description 1
- 230000031068 symbiosis, encompassing mutualism through parasitism Effects 0.000 description 1
- 239000002699 waste material Substances 0.000 description 1
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Classifications
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- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03B—SEPARATING SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS
- B03B7/00—Combinations of wet processes or apparatus with other processes or apparatus, e.g. for dressing ores or garbage
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/001—Flotation agents
- B03D1/004—Organic compounds
- B03D1/008—Organic compounds containing oxygen
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/001—Flotation agents
- B03D1/004—Organic compounds
- B03D1/012—Organic compounds containing sulfur
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D1/00—Flotation
- B03D1/001—Flotation agents
- B03D1/018—Mixtures of inorganic and organic compounds
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D2201/00—Specified effects produced by the flotation agents
- B03D2201/02—Collectors
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- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D2201/00—Specified effects produced by the flotation agents
- B03D2201/04—Frothers
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- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D2201/00—Specified effects produced by the flotation agents
- B03D2201/06—Depressants
-
- B—PERFORMING OPERATIONS; TRANSPORTING
- B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
- B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
- B03D2203/00—Specified materials treated by the flotation agents; Specified applications
- B03D2203/02—Ores
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- Chemical & Material Sciences (AREA)
- Inorganic Chemistry (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention discloses a process method for recovering sulfur concentrate from potassium-rich slate, which comprises the steps of crushing, grinding and magnetic separation of the potassium-rich slate, and then adding an activating agent, a collecting agent and a foaming agent for rough concentration and fine concentration to obtain sulfur concentrate; wherein the magnetic field intensity of the magnetic separation is 159-199kA/m, the activating agent is a mixture of oxalic acid and copper sulfate, the collecting agent is a mixture of isobutyl xanthate and mixed amine, and the foaming agent is pine oil. The process method for recovering the sulfur concentrate from the potassium-rich slate optimizes the combination ratio of the flotation reagents, effectively improves the separation efficiency of the sulfur-containing concentrate, obtains the high-quality sulfur concentrate, saves the cost and improves the income.
Description
Technical Field
The invention relates to the technical field of mineral engineering, in particular to a process method for recovering sulfur concentrate.
Background
The Baiyunebo ore main ore and the Dong ore contain rich potassium-rich slate resources, the reserve of the potassium-rich slate is proved to be 3.4 hundred million tons in the boundary of the upper pan of the main ore and the Dong ore, and the Baiyunebo ore belongs to an oversized potassium ore deposit. The potassium-rich slate ore mainly comprises potash feldspar, and then comprises amphibole, biotite, sodalite, pyrrhotite, hematite and the like, and a small amount of quartz, pyrite and the like. Wherein pyrrhotite and hematite are in irregular flaky symbiosis, and the local content can reach 10%. Generally, the average grade of potassium oxide of a potassium ore is 10.58%, the sulfur content is about 2.4%, if the ore is directly roasted, as the grade of potassium oxide is low, the sulfur content is higher, more slag amount and adverse effect are generated in the roasting process, therefore, before the ore is utilized, the ore is subjected to ore dressing, gangue minerals such as pyrrhotite, pyrite, sodalite, hornblende, biotite and the like are removed, qualified potash feldspar ore concentrate is selected, the sulfur is recovered, the subsequent effect is reduced, and the benefit is improved.
In recent years, a plurality of domestic scientific research units carry out a great deal of research on the comprehensive utilization and research of potassium-rich slate ores and obtain stage results, but research on the aspects of direct roasting, leaching and the like of potassium-rich slate as a raw material is carried out on the research raw material. The potassium-rich slate is not subjected to mineral separation and how to select qualified potash feldspar for research, so that a large amount of slag is generated in subsequent research, and the subsequent operations such as roasting, leaching and the like are also adversely affected due to the existence of sulfur.
Pyrrhotite is easy to oxidize and argillize and belongs to a mineral which is difficult to select. At present, more research is carried out on ore dressing for removing pyrrhotite, but the ore dressing for recovering potash feldspar ore concentrate and sulfur ore concentrate from potassium-rich slate is rare. At present, the mineral processing technology for recovering potash feldspar ore concentrate mainly adopts a mineral processing technology for removing iron and biotite, and the flotation method is generally adopted for recovering pyrrhotite, so that the research on flotation inhibitors and activators is more. Therefore, when the qualified potash feldspar ore concentrate is selected, the superior product sulfur ore concentrate is selected, so that the requirements of further comprehensively recycling the sulfur-containing minerals and changing waste into valuables are met.
Therefore, the technical personnel in the field are dedicated to develop a process method for recovering sulfur concentrate from the potassium-rich slate, and the defects of high sulfur content, high subsequent roasting slag amount, sulfur concentrate loss and the like of the potassium-rich slate in the prior art are overcome.
Disclosure of Invention
In view of the above defects of the prior art, the technical problems to be solved by the invention are that the sulfur content of the potassium-rich slate ore is low, the roasting slag amount is high, and the sulfur concentrate is lost when the sulfur concentrate is recovered.
In order to achieve the purpose, the invention provides a process method for recovering sulfur concentrate from potassium-rich slates, which comprises the steps of crushing, grinding and magnetic separation of the potassium-rich slates, and then adding an activating agent, a collecting agent and a foaming agent for roughing and fine separation to obtain sulfur concentrate; wherein the magnetic field intensity of the magnetic separation is 159-199kA/m, the activating agent is a mixture of oxalic acid and copper sulfate, the collecting agent is a mixture of isobutyl xanthate and mixed amine, and the foaming agent is pine oil.
Further, in the mixture of oxalic acid and copper sulfate, the weight (gram) ratio of the oxalic acid to the copper sulfate is 3:1-5: 1;
further, in the mixture of the isobutyl xanthate and the mixed amine, the weight (gram) ratio of the isobutyl xanthate to the mixed amine is 5:1-7: 1.
Further, the process method for recovering the potassium feldspar in the potassium-rich slate comprises the following steps:
step 1, crushing potassium-rich slates;
step 2, grinding the potassium-rich slate crushed in the step 1;
step 3, carrying out magnetic separation on the ground ore product obtained in the step 2 to obtain magnetic concentrate and magnetic tailings;
and 4, adding an activating agent and a collecting agent into the magnetic separation tailings obtained in the step 3 for flotation and sorting to obtain flotation foam and potash feldspar ore concentrate.
And 5, mixing the magnetic concentrate obtained in the step 3 with the flotation foam obtained in the step 4, and adding an activating agent, a collecting agent and a foaming agent to perform rough concentration and fine concentration to obtain the sulfur concentrate.
Further, in the step 1, crushing the potassium-rich slate into three sections of coarse crushing, medium crushing and fine crushing, wherein the granularity of a crushed product is 0-3 mm;
further, in the step 2, in the ore grinding product, the product with the granularity of 0-74um accounts for 70-85% of the total amount;
further, in the step 3, the magnetic field intensity of the magnetic separation is 159-;
further, in the step 4, the activating agent is a mixture of oxalic acid and copper sulfate; the collector is a mixture of isobutyl xanthate and mixed amine;
further, in the step 4, the dosage of the activating agent is 1.0-1.6kg/t during flotation and sorting; the dosage of the collecting agent is 0.05-0.09 kg/t;
further, in the step 4, the pH value of the flotation is 5.5-6.0;
further, in the step 4, the temperature for flotation separation is 15-30 ℃;
further, in the step 4, the activating agent is a mixture of oxalic acid and copper sulfate, wherein the weight (gram) ratio of the oxalic acid to the copper sulfate is 2: 1;
further, in the step 4, the collector is a mixture of isobutyl xanthate and mixed amine, wherein the weight (gram) ratio of the isobutyl xanthate to the mixed amine is 4: 1.
Further, in the step 5, the activating agent is a mixture of oxalic acid and copper sulfate, wherein the weight (gram) ratio of the oxalic acid to the copper sulfate is 3:1-5: 1;
further, in the step 5, the collecting agent is a mixture of isobutyl xanthate and amine alcohol xanthate, wherein the weight (gram) ratio of the isobutyl xanthate to the amine alcohol xanthate is 5:1-7: 1;
further, in the step 5, the foaming agent is pine oil;
further, in the step 5, during rough concentration, the dosage of the activating agent is 1.0-1.5kg/t, the dosage of the collecting agent is 0.05-0.08kg/t, and the dosage of the foaming agent is 0.03-0.05 kg/t; the rough concentration PH value is 5.0-6.0;
further, in the step 5, the concentration comprises three times of concentration, wherein the dosage of the collecting agent in the first concentration is 0.03-0.05kg/t, and the dosage of the foaming agent is 0.01-0.02 kg/t; the dosage of the collecting agent for the second time of concentration is 0.01-0.02kg/t, and the dosage of the foaming agent is 0.01-0.02 kg/t; the dosage of the collecting agent for the third time of concentration is 0.01-0.02kg/t, and the dosage of the foaming agent is 0.01-0.02 kg/t.
In the preferred embodiment of the invention, in the step 3, the magnetic field intensity for magnetic separation is 179 kA/m; (ii) a
In a preferred embodiment of the invention, in the step 4, the amount of the activating agent used in the flotation is 1.2 kg/t;
in another preferred embodiment of the present invention, in the step 4, the amount of the activating agent used in the flotation is 1.0 kg/t;
in another preferred embodiment of the present invention, in the step 4, the amount of the activating agent used in the flotation is 1.6 kg/t;
in a preferred embodiment of the invention, in the step 4, the amount of the collecting agent is 0.07kg/t during flotation and sorting;
in another preferred embodiment of the invention, in the step 4, the amount of the collector used in the flotation is 0.05 kg/t;
in another preferred embodiment of the invention, in the step 4, the amount of the collector used in the flotation is 0.09 kg/t;
in a preferred embodiment of the present invention, in the step 4, the PH value of the flotation separation is 5.5;
in another preferred embodiment of the present invention, in the step 4, the PH value of the flotation separation is 6.0;
in a preferred embodiment of the present invention, in the step 5, the activating agent is a mixture of oxalic acid and copper sulfate, wherein the weight (g) ratio of oxalic acid to copper sulfate is 4: 1;
in another preferred embodiment of the present invention, in the step 5, the activating agent is a mixture of oxalic acid and copper sulfate, wherein the weight (g) ratio of oxalic acid to copper sulfate is 3: 1;
in another preferred embodiment of the present invention, in the step 5, the activating agent is a mixture of oxalic acid and copper sulfate, wherein the weight (g) ratio of oxalic acid to copper sulfate is 5: 1;
in a preferred embodiment of the present invention, in the step 5, the collector is a mixture of isobutyl xanthate and amine alcohol xanthate, wherein the weight (g) ratio of isobutyl xanthate to amine alcohol xanthate is 6: 1;
in another preferred embodiment of the present invention, in the step 5, the collector is a mixture of isobutyl xanthate and amine alcohol xanthate, wherein the weight (g) ratio of isobutyl xanthate to amine alcohol xanthate is 5: 1;
in another preferred embodiment of the present invention, in the step 5, the collector is a mixture of isobutyl xanthate and amine alcohol xanthate, wherein the weight (g) ratio of isobutyl xanthate to amine alcohol xanthate is 7: 1;
in a preferred embodiment of the present invention, in the step 5, the amount of the activating agent used in the roughing is 1.2 kg/t;
in a preferred embodiment of the invention, in the step 5, the amount of the collector is 0.07 kg/t;
in a preferred embodiment of the present invention, in the step 5, the amount of the foaming agent is 0.04 kg/t;
in a preferred embodiment of the present invention, in the step 5, the rougher PH is 5.2;
in the preferred embodiment of the invention, the concentration in the step 5 comprises three times of concentration, wherein the dosage of the collecting agent in the first concentration is 0.035kg/t, and the dosage of the foaming agent is 0.015 kg/t; the dosage of the collecting agent for the second time of concentration is 0.02kg/t, and the dosage of the foaming agent is 0.01 kg/t; the dosage of the collecting agent for the third time of concentration is 0.01kg/t, and the dosage of the foaming agent is 0.01 kg/t.
By adopting the scheme, the process method for recovering the sulfur concentrate from the potassium-rich slate disclosed by the invention has the following technical effects:
(1) the process method adopts reasonable and feasible mineral separation process and the combination of collecting property and selective and efficient flotation reagent, effectively recovers the sulfur concentrate from the potassium-rich slate, reduces the impurities and harmful element sulfur in the potassium feldspar, improves the content of potassium oxide and reduces the sulfur content in the potassium feldspar, obtains qualified potassium feldspar concentrate, reduces the slag amount of subsequent operation and the adverse effect of sulfur, changes the harmful sulfur-containing minerals into valuables, obtains the selection of superior sulfur concentrate, improves the utilization rate and realizes high yield;
(2) the process method of the invention obtains the high-quality potash feldspar ore concentrate, simultaneously takes the mixed magnetic separation ore concentrate and flotation foam as the raw material for flotation of the sulfur-containing minerals, adopts simple and efficient flotation reagent combination to separate the content minerals through flotation operation, and obtains the superior sulfur ore concentrate.
In conclusion, the process method for recovering the sulfur concentrate from the potassium-rich slate optimizes the combination ratio of the flotation reagents, effectively improves the separation efficiency of the sulfur-containing concentrate, obtains the high-quality sulfur concentrate, saves the cost and improves the benefit.
The conception, the specific structure and the technical effects of the present invention will be further described with reference to the accompanying drawings to fully understand the objects, the features and the effects of the present invention.
Drawings
FIG. 1 is a process flow diagram of an embodiment of the invention;
Detailed Description
The technical contents of the preferred embodiments of the present invention will be more clearly and easily understood by referring to the drawings attached to the specification. The present invention may be embodied in many different forms of embodiments and the scope of the invention is not limited to the embodiments set forth herein.
Example, process method for recovering sulfur concentrate from potassium-rich slate
Step 1, treatment of raw materials
Carrying out three-stage crushing of coarse crushing, medium crushing and fine crushing on the potassium-rich slate, wherein the granularity of a crushed product is-3 mm;
step 2, grinding ore
Grinding the crushed product, wherein the granularity of the ground product is-0.074 mm and accounts for 81.4%;
step 3, carrying out primary magnetic separation on the ground ore product under the condition of magnetic field intensity of 179.1kA/m to obtain magnetic concentrate and magnetic tailings;
through detection, the sulfur content in the magnetic concentrate is 10.41 percent, and the recovery rate is 44.50 percent; the content of potassium oxide in the magnetic separation tailings is 10.70 percent, the recovery rate of potassium oxide is 95.33 percent, and the sulfur content is 1.41 percent;
step 4, using the magnetic separation tailings obtained in the step as raw materials of sulfide ore, biotite and tremolite, carrying out flotation separation on the magnetic separation tailings at normal temperature, using oxalic acid and copper sulfate in a mass ratio of 2:1 as activating agents of sulfur-containing minerals, using mixed amine in a mass ratio of 4:1 as collecting agents and foaming agents of the minerals such as sulfide ore, biotite and tremolite, firstly adding 1.2kg/t of the activating agents, stirring for 5min, adjusting the pH value to 5.5, then adding 0.07kg/t of the collecting agents and foaming agents, stirring for 2min, carrying out rough separation, and carrying out rough separation for 6min to obtain potassium feldspar concentrate (the recovery rate of potassium oxide is 82.02%) with the potassium oxide content of 12.15%, the sulfur-containing content of 0.18% and flotation foam with the sulfur content of 5.28%;
step 5, mixing the magnetic concentrate obtained in the step and flotation foam to obtain a mixture serving as a raw material for flotation of sulfur-containing minerals, adding 1.2kg/t of activating agent at the natural ore pulp temperature by using oxalic acid, copper sulfate and isobutyl xanthate and amine alcohol xanthate in a mass ratio of 4:1 as activating agents for the sulfur-containing minerals, adding 0.07kg/t of collecting agent after stirring for 2min, adding 0.04kg/t of foaming agent after stirring for 2min, and performing rough separation after stirring for 1min to obtain sulfur rough concentrate and rough separation tailings; adding 0.35kg/t of collecting agent at the pH value and the temperature of natural ore pulp, stirring for 2min, adding 0.015kg/t of foaming agent, stirring for 1min, and carrying out primary concentration on the sulfur roughing concentrate to obtain a sulfur concentrate 1 and a middling 1; adding 0.02kg/t of collecting agent at the pH value and the temperature of natural ore pulp, stirring for 2min, adding 0.01kg/t of foaming agent, stirring for 1min, and performing secondary concentration on the sulfur concentrate 1 to obtain a sulfur concentrate 2 and a middling 2; adding 0.01kg/t of collecting agent at the pH value and the temperature of natural ore pulp, stirring for 2min, adding 0.01kg/t of foaming agent, stirring for 1min, and then carrying out three-time concentration on the sulfur concentrate 2 to obtain sulfur concentrate with 39.00% of sulfur content and 47.03% of recovery rate and middling 3;
step 6, returning middlings 1, middlings 2 and middlings 3 generated in the sulfide flotation operation to the primary sulfur roughing operation, the primary concentration operation and the secondary concentrate operation in sequence, wherein the roughed tailings are used as final tailings;
by adopting the ore grinding-magnetic separation-mixed flotation-flotation sulfur separation process, potassium feldspar ore concentrate with the potassium oxide content of 12.15 percent and the sulfur content of 0.18 percent (the potassium oxide recovery rate is 82.02 percent) and sulfur ore concentrate with the sulfur content of 39.00 percent (the sulfur recovery rate is 47.03 percent) can be obtained from potassium-rich slate containing 10.12 percent of potassium oxide and 2.30 percent of sulfur. The sulfur content and the recovery rate of the obtained sulfur concentrate are high, the sulfur concentrate in the potassium-rich slate is effectively recovered, the content of sulfur in potassium feldspar is reduced, and the method is beneficial to protecting the atmospheric environment.
Other technical schemes of the invention also have similar using effects.
The foregoing detailed description of the preferred embodiments of the invention has been presented. It should be understood that numerous modifications and variations could be devised by those skilled in the art in light of the present teachings without departing from the inventive concepts. Therefore, the technical solutions available to those skilled in the art through logic analysis, reasoning and limited experiments based on the prior art according to the concept of the present invention should be within the scope of protection defined by the claims.
Claims (2)
1. A process method for recovering sulfur concentrate from potassium-rich slates is characterized by comprising the steps of crushing, grinding and magnetic separation of the potassium-rich slates, and then adding an activating agent, a collecting agent and a foaming agent for rough concentration and fine concentration to obtain sulfur concentrate; wherein the magnetic field intensity of the magnetic separation is 159-;
the process method comprises the following specific steps:
step 1, crushing potassium-rich slates;
step 2, grinding the potassium-rich slate crushed in the step 1;
step 3, carrying out magnetic separation on the ground ore product obtained in the step 2 to obtain magnetic concentrate and magnetic tailings;
and 4, adding an activating agent and a collecting agent into the magnetic separation tailings obtained in the step 3 for flotation and sorting to obtain flotation foam and potash feldspar ore concentrate.
Step 5, mixing the magnetic concentrate obtained in the step 3 with the flotation foam obtained in the step 4, and adding an activating agent, a collecting agent and a foaming agent to perform rough concentration and fine concentration to obtain sulfur concentrate;
in the step 2, the product with the granularity of-74 um accounts for 70-85% of the total amount in the ore grinding product;
in the step 4, the activating agent is a mixture of oxalic acid and copper sulfate, and the weight ratio of the oxalic acid to the copper sulfate is 2: 1; the collecting agent is a mixture of isobutyl xanthate and mixed amine, and the weight ratio of the isobutyl xanthate to the mixed amine is 4: 1; during flotation and sorting, the dosage of the activating agent is 1.0-1.6 kg/t; the dosage of the collecting agent is 0.05-0.09 kg/t; the pH value of the flotation is 5.5-6.0; the flotation temperature is 15-30 ℃;
in the step 5, the activating agent is a mixture of oxalic acid and copper sulfate, and the weight ratio of the oxalic acid to the copper sulfate is 4: 1; the collecting agent is a mixture of isobutyl xanthate and amine alcohol xanthate, and the weight ratio of the isobutyl xanthate to the amine alcohol xanthate is 6: 1;
in the step 5, the dosage of the activating agent is 1.0-1.5kg/t, the dosage of the collecting agent is 0.05-0.08kg/t and the dosage of the foaming agent is 0.03-0.05kg/t during rough concentration; the rough concentration PH value is 5.0-6.0;
in the step 5, the concentration comprises three times of concentration, wherein the dosage of the collecting agent in the first concentration is 0.03-0.05kg/t, and the dosage of the foaming agent is 0.01-0.02 kg/t; the dosage of the collecting agent for the second time of concentration is 0.01-0.02kg/t, and the dosage of the foaming agent is 0.01-0.02 kg/t; the dosage of the collecting agent for the third time of concentration is 0.01-0.02kg/t, and the dosage of the foaming agent is 0.01-0.02 kg/t.
2. The process method as claimed in claim 1, wherein in the step 1, the potassium-rich slate is crushed into three sections of coarse crushing, medium crushing and fine crushing, and the particle size of the crushed product is-3 mm.
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| CN102284351B (en) * | 2011-07-28 | 2013-08-28 | 内蒙古科技大学 | Beneficiation method for separating potassium feldspar concentrates from BayanObo potassium-rich slates |
| CN102357402B (en) * | 2011-07-28 | 2013-05-08 | 内蒙古科技大学 | Method for separating potassium feldspar ore concentrate and iron ore concentrate from Baiyunebo potassium rich slate |
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