CN108929947A - A method of comprehensive utilization kaliophilite mine - Google Patents

A method of comprehensive utilization kaliophilite mine Download PDF

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CN108929947A
CN108929947A CN201810936659.6A CN201810936659A CN108929947A CN 108929947 A CN108929947 A CN 108929947A CN 201810936659 A CN201810936659 A CN 201810936659A CN 108929947 A CN108929947 A CN 108929947A
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kaliophilite
mine
solution
sodium
slag
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CN108929947B (en
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申晓毅
邵鸿媚
刘佳囡
翟玉春
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Northeastern University China
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    • C05—FERTILISERS; MANUFACTURE THEREOF
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    • C05G—MIXTURES OF FERTILISERS COVERED INDIVIDUALLY BY DIFFERENT SUBCLASSES OF CLASS C05; MIXTURES OF ONE OR MORE FERTILISERS WITH MATERIALS NOT HAVING A SPECIFIC FERTILISING ACTIVITY, e.g. PESTICIDES, SOIL-CONDITIONERS, WETTING AGENTS; FERTILISERS CHARACTERISED BY THEIR FORM
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    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
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    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
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    • C22B26/10—Obtaining alkali metals
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    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
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    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
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    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
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    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
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    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/44—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes
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Abstract

The present invention relates to a kind of methods for comprehensively utilizing kaliophilite mine, include the following steps: S1, pre-process to kaliophilite mine；S2, pretreated kaliophilite mine is uniformly mixed with sodium carbonate, then carries out calcination process, obtains the first roasting material.S3, the first roasting material water or sodium hydroxide solution are dissolved out, is then filtered separation, obtains the pre-desiliconizing slag containing potassium and aluminium.S4, pre-desiliconizing slag is uniformly mixed with ammonium sulfate, is then roasted, obtain the second roasting material.S5, by second roasting material with it is water-soluble out, be then filtered separation, obtain silicon slag and the aluminum ammonium sulfate solution containing potassium.The method that the present invention comprehensively utilizes kaliophilite mine can greatly improve element utilization rate in kaliophilite mine, realize few high value added utilization, waste residue yield, simple process and be suitble to large-scale production.

Description

A method of comprehensive utilization kaliophilite mine
Technical field
The invention belongs to non-ferrous metallurgy technology fields, and in particular to a method of comprehensive utilization kaliophilite mine.
Background technique
China is aluminum i ndustry big country, and aluminium yield accounts for about the 40% of world aluminum yield, but China can be used for industrial aluminium Native ore resources reserves are low, and the exploitation of bauxite is not high using technology content, leads to serious waste of resources, so that the alumina in China Ore resources degree of protection further decreases.The discovery of the high-iron bauxites, high-sulfur bauxite such as Fujian Zhangpu, Guangxi Guigang, greatly The status for alleviating China's bauxite scarcity, the characteristics of because of its high-speed rail, high silicon, high-sulfur, although there is different Utilization plans, But rational exploitation and utilization is not yet received.
China is also large agricultural country, and crop yield position is at the forefront in the world, but with the lasting improvement of agricultural production conditions The increase of demand with soil to valuable element, potash fertilizer demand increasingly increase.According to the investigation of the Chinese Academy of Agricultural Sciences, the shortage of potassium is had become For the main restricting factor of many regional increasing crop yields.However, the soluble potassium critical shortage in China, the world reserves Jin Zhan are total The 0.63% of reserves, and be distributed extremely uneven.China's non-solubility potassium resource reserves are abundant, and many kinds of, such as potassium feldspar, potassium Nepheline, Potassium-rich shale, illite etc., distribution is almost throughout the whole nation.
China's nepheline ore resources are abundant, and explored reserves in 10,000,000,000 t or more, are distributed in the multiple provinces and regions in the whole nation. China's nepheline ore resources potassium content, which is above 10%, aluminium content, also can reach 20% or so, and be rich in silicon resource, sodium, iron content It is low, great Development volue.Therefore, feasible technology path is developed, realizes the synthesis of valuable element potassium, aluminium, silicon, iron in nepheline mine Extraction and application is both to solve the effective way of China's potash fertilizer shortage status, and alleviate the effective of China's bauxite resource shortage Approach, or improve the effective way of the level of resources utilization.
Currently, nepheline mine, which produces sylvite and the treatment process of aluminium oxide, mainly to be had: soda lime sintering process, high pressure aquation method, Acid system etc..Soda lime sintering process can use the production of the useful constituent in mineral aluminium oxide, the products such as potassium carbonate, but after mentioning aluminium Residue red mud amount is big, can be used for producing cement and siliceous fertilizer, but limited by cement sale radius and the hardened soil of siliceous fertilizer, silicon utilization rate It is lower.High pressure aquation method can decompose the potassium and aluminium extracted in kaliophilite in lower temperature, and leached mud is mainly calcium silicates, can water processed Mud, but the requirement of this method technique is stringent, mass flow is big, and tailings amount is big.Acid hydrolyzation can decompose kaliophilite and extract potassium therein And aluminium, it is stayed in slag through divided silicon, then extraction and application, but contain potassium feldspar in kaliophilite mine, acid system is not easily decomposed, and reduces potassium With the utilization rate of aluminium.And the common problem of the above method is to be conceived to the potassium and aluminium extracted in nepheline, the utilization rate of silicon is low, thus A large amount of waste residues are generated, are chiefly used in producing the low cement of added value, environmental pressure is big.
Therefore, the new process for exploring a comprehensive utilization kaliophilite mine realizes that the greenization of kaliophilite ore resources, height are additional Value comprehensive utilization, meets the needs of economizing on resources and is environmental-friendly, has important practical significance and application value.
Summary of the invention
(1) technical problems to be solved
In order to solve the above problem of the prior art, the present invention, which provides one kind, can greatly improve element benefit in kaliophilite mine With rate, realize few high value added utilization, waste residue yield, simple process and the comprehensive utilization kaliophilite mine for being suitble to large-scale production Method.
(2) technical solution
In order to achieve the above object, the main technical schemes that the present invention uses include:
The present invention provides a kind of method for comprehensively utilizing kaliophilite mine, includes the following steps: S1, carries out in advance to kaliophilite mine Processing；S2, pretreated kaliophilite mine is uniformly mixed with sodium carbonate, then carries out calcination process, obtains the first calcining matter Material, maturing temperature are 800~1000 DEG C, and the soaking time of roasting is 50~150min；S3, by first roasting material water or Sodium hydroxide solution dissolution, is then filtered separation, obtains the pre-desiliconizing slag containing potassium and aluminium；S4, by pre-desiliconizing slag and sulfuric acid Ammonium is uniformly mixed, and is then roasted, and obtains the second roasting material, and maturing temperature is 350~550 DEG C, and calcining time is 50~ 150min；S5, then the second roasting material is filtered separation out with water-soluble, obtains silicon slag and alum containing potassium is molten Liquid；It being added in S6, the aluminum ammonium sulfate solution containing potassium obtained to step S5 and contains ammonium salt solution, the pH value for adjusting solution carries out heavy aluminium, Then it is filtered separation, obtains the mixed solution of thick aluminum hydroxide precipitation and potassium sulfate and ammonium sulfate；S7, step S6 are obtained Potassium sulfate and the mixed solution of ammonium sulfate fractional crystallization preparing potassium sulfate crystal and ammonia sulfate crystal are carried out after circulation collection, S4 is as raw material for ammonia sulfate crystal return step；Or the mixed solution of the obtained potassium sulfate of step S6 and ammonium sulfate is directly made For composite chemical fertilizer use.
Wherein, the purpose roasted in step s 2 is mainly by KAlSi3O8Structure carry out destroy make its decomposition, work as roasting When temperature is less than 800 DEG C, KAlSi3O8Be difficult to decompose, and temperature be greater than 1000 DEG C when, consume energy it is higher.Therefore, maturing temperature is best Control can guarantee that energy consumption is lesser while KAlSi at 800~1000 DEG C3O8Decomposition can be normally destroyed, pre-desiliconizing is improved Recovery rate.The soaking time of roasting is preferably controlled in 50~150min, so that calcination is more abundant.Maturing temperature is into one Preferably 850 DEG C of step, the soaking time of roasting is 90min, and with this condition, the recovery rate of pre-desiliconizing is best.
Further, the purpose that roasts again in step s 4 primarily to roast step S2 after obtain KAlSiO4Further decompose, when maturing temperature is less than 350 DEG C, be difficult to decompose, and temperature be greater than 550 DEG C when, consume energy it is higher. Therefore, maturing temperature is preferably controlled in 350~550 DEG C, can guarantee that energy consumption is lesser while KAlSiO4It can normally decompose, Improve the dissolution rate of potassium aluminium.The soaking time of roasting is preferably controlled in 50~150min, so that calcination is more abundant.Roasting Burning temperature is more preferably 475 DEG C, and the soaking time of roasting is 75min, with this condition, improves the dissolution rate of potassium aluminium most It is good.
A preferred embodiment according to the present invention, in step sl, pretreatment include being crushed kaliophilite mine, then by potassium The granularity of nepheline mine is levigate to 80 μm or less；In step s 2, the dosage of sodium carbonate is that silica is just complete in kaliophilite mine 0.8~1.4 times of sodium carbonate theoretical amount required for full response, is roasted using mechanical activation mode；In step s3, it uses The concentration of sodium hydroxide solution used is 0.01~0.4mol/L when sodium hydroxide solution is dissolved out, solid-liquid quality when dissolution Than for 1:2~6, leaching temperature is 50~95 DEG C, dissolution time is 20~120min.
Wherein, the chemical component of kaliophilite mine used specifically includes that SiO by weight percentage240~70%, Al2O3 20~35%, K2O 10~24%, Na2O 0~1.0% and Fe2O31~4%.
Further, it is further preferably dissolved out in step s3 using sodium hydroxide solution, can further shorten dissolution Time.The effect of sodium hydroxide solution is primarily to make the Na in the first roasting material in step S32SiO3Dissolution, while energy A small amount of amorphous SiO enough and that may be present2Reaction.Therefore, the sodium hydroxide solution in this step selects lower solubility 0.01~0.4mol/L.Under this leaching temperature, amorphous SiO can be made2It is sufficiently reacted with NaOH solution.Further preferably It is 85 DEG C for leaching temperature, dissolution time 90min can guarantee that energy consumption is smaller while dissolution rate is very fast.
A preferred embodiment according to the present invention, in step s 4, the dosage of ammonium sulfate are potassium and aluminium in pre-desiliconizing slag Just 1.0~1.6 times of ammonium sulfate theoretical amount required for reacting completely；In step s 5, the solid-liquid of the second roasting material and water Mass ratio is 1:2~6, and leaching temperature is 60~90 DEG C, and dissolution time is 30~120min；In step s 6, it is containing ammonium salt solution At least one of ammonium hydroxide and ammonium hydrogen carbonate, pH value are 5.4~5.6.
Wherein, in step s 6, in the range of this pH value, aluminium can be made to precipitate completely, and potassium not will form precipitating, And then separate aluminium and potassium.
A preferred embodiment according to the present invention further includes following steps after step S6: S8, thick aluminium hydroxide sinking Shallow lake is dissolved out with aqueous slkali, is then filtered separation, is obtained sodium aluminate solution and tailings；S9, sodium aluminate solution is subjected to carbon point, Then it is filtered separation, obtains aluminium hydroxide and sodium carbonate liquor.
A preferred embodiment according to the present invention, further includes following steps after step s 5: S10, obtaining step S5 Silicon slag leached with sodium hydroxide solution, then to leach slurry be diluted, be then filtered separation, obtain sodium silicate solution And filter residue；S11, the sodium silicate solution for obtaining step S10 carry out first time carbon point, are then filtered separation, obtain silicic acid The mixed liquor and miscellaneous slag of sodium and sodium carbonate；S12, the mixed liquor of the obtained sodium metasilicate of step S11 and sodium carbonate is carried out second Secondary carbon point, is then filtered separation, obtains sodium carbonate liquor and white carbon black.
A preferred embodiment according to the present invention, in step s 8, aqueous slkali are sodium hydroxide solution, sodium hydroxide solution Concentration be 0.5~2mol/L, leaching temperature be 70~95 DEG C；In step slo, the concentration of sodium hydroxide solution be 4~ The solid-liquid mass ratio of 15mol/L, silicon slag and sodium hydroxide solution is 1:1.5~5, and extraction temperature is 75~145 DEG C, extraction time For 30~90min, the concentration for leaching sodium metasilicate in solution after pulp dilution is 1~2.5mol/L；In step s 11, for the first time Carbon divides the pH value of terminal to be 11.0~11.5；In step s 12, second of carbon divides the pH value of terminal to be 8.5~9.
Wherein, in step s 8, the effect of sodium hydroxide solution is primarily to dissolve out thick aluminium hydroxide, due to dissolving out In journey reaction speed quickly, therefore in step s 8 sodium hydroxide solution select 0.5~2mol/L of low concentration.
Further, in step slo, the selection of extraction temperature is related with the concentration selection of sodium hydroxide solution.Due to The solubility of itself of sodium hydroxide is very big, and when the concentration of sodium hydroxide solution is greater than 15mol/L, alkalinity is too strong, on the one hand General filter paper and filter cloth are difficult to bear when being filtered, on the other hand subsequent also to need to be added more water and be diluted, and When concentration is less than 4mol/L, inventory is larger, and therefore, the concentration of sodium hydroxide solution is preferably controlled in 4~15mol/L.Herein Under concentration range, extraction temperature is preferably controlled in 75~145 DEG C, can make to leach more abundant.
Further, first time carbon divides the pH value of terminal to control 11.0~11.5, can have the function that removal of impurities and Guarantee the quality of sodium silicate solution the carbon mitogenetic yield for producing white carbon black and white carbon black.And second of carbon divides the pH value of terminal to control 8.5~9, it can guarantee that the sodium metasilicate in solution can sufficiently react.
A preferred embodiment according to the present invention, in step s 2, the tail gas that calcination process generates are molten using sodium hydroxide Liquid absorbs, and return step S2 is as raw material after then evaporative crystallization obtains crystals of sodium carbonate；Or to the tail that calcination process generates Gas is collected, for carbon point in step S9, step S11 and step S12；Or the tail gas that calcination process generates uses ammonia solution It absorbs, is used to sink aluminium in step S6 after obtaining ammonium hydrogen carbonate；In step s 4, the flue gas roasted is absorbed using dilute sulfuric acid, Obtained ammonium sulfate again after inspissation crystallizes return step S4 as raw material；In step s 6, using bicarbonate ammonia-sinking aluminium The gas generated in the process is absorbed using sodium hydroxide solution, the obtained sodium carbonate liquor return step S2 after inspissation crystallizes again As raw material；In step s 11, the miscellaneous slag obtained in return step S10 with silicon slag for mixing together.
A preferred embodiment according to the present invention, further includes following steps after step s 12: S13, to sodium carbonate liquor Middle addition active lime lotion carries out causticizing reaction, is then filtered separation, obtains calcium carbonate and sodium hydroxide solution；S14, The sodium hydroxide solution that step S13 is obtained carry out inspissation to a certain concentration, and obtained sodium hydroxide inspissation solution returns Silicon slag is leached in step S10.
A preferred embodiment according to the present invention, in step s 13, the molar ratio of active lime cream and sodium carbonate are 0.9 ~1.2:1, causticization temperature are 50~95 DEG C, and the causticization time is 10~90min, and the concentration of active lime lotion is 140~240g/ L。
Wherein, when causticization temperature is less than 50 DEG C, causticizing efficiency is too low, and production efficiency is too low.Therefore, causticization temperature preferably controls At 50~95 DEG C, causticizing efficiency, throughput rate and production efficiency are best.
A preferred embodiment according to the present invention, in step s 9, obtained sodium carbonate liquor are returned after inspissation crystallizes again Step S2 is returned as raw material or for causticization in step S13；In step s 12, the sodium carbonate liquor obtained is again through inspissation knot Use or be used for causticization in step S13 after crystalline substance in return step S2 as raw material；In step S14, the steaming that is obtained after inspissation Distilled water is for recycling.
(3) beneficial effect
The beneficial effects of the present invention are:
Kaliophilite mine is sufficiently decomposed by the way of double roasting in the present invention, first carries out first using sodium carbonate Secondary roasting, and control maturing temperature at 800~1000 DEG C, it can be by KAlSi in kaliophilite mine3O8Structure carry out destroy make it Decomposition obtains KAlSiO4, realize pre-desiliconizing, so make originally very stable kaliophilite mine become active, it is more conducively subsequent with Ammonium sulfate is reacted.Then second is carried out using ammonium sulfate again to roast, and control maturing temperature at 350~550 DEG C, it can By KAlSiO4Further reaction decomposes and obtains K2SO4、NH4Al(SO4)2And SiO2, realize that potassium aluminium extracts, and then by kaliophilite mine In potassium, aluminium and silicon be sufficiently separated.By the way that the structure of kaliophilite mine can be sufficiently destroyed after double roasting, make potassium therein, Aluminium and silicon are sufficiently separated, and are more advantageous to subsequent processing and are obtained the products such as potassium sulfate, substantially increase to the element in kaliophilite mine Utilization rate, realize high value added utilization.And the by-product obtained during processing can be recycled recycling, section About resource, also improves the utilization rate of resource.
In addition, using sodium carbonate roasting and ammonium sulfate roasting double roasting carry out decompose kaliophilite mine method, energy consumption compared with Low, the waste residue of generation is less, and the waste residue and tail gas that generate can recycle after processing, substantially increase the weight of resource It is multiple to utilize, meet the needs of economizing on resources and is environmental-friendly.Preparation method simple process simultaneously is suitble to large-scale production.
Detailed description of the invention
Fig. 1 is the method flow diagram that kaliophilite mine is comprehensively utilized in the present invention.
Specific embodiment
In order to preferably explain the present invention, in order to understand, with reference to the accompanying drawing, by specific embodiment, to this hair It is bright to be described in detail.
Embodiment 1
The present embodiment provides a kind of methods for comprehensively utilizing kaliophilite mine, specifically comprise the following steps:
S1, pretreatment: kaliophilite mine is pre-processed.
Specifically, the chemical composition of kaliophilite mine used in the present embodiment specifically includes that SiO by weight percentage2 62.35%, Al2O320.78%, K2O 12.66%, Na2O 0.70% and Fe2O33.16%.Pretreatment mainly includes first will Kaliophilite mine is broken, then that the granularity of kaliophilite mine is levigate to 80 μm hereinafter, so that subsequent reactions are more abundant.
S2, mixing and roasting pre-desiliconizing: by pretreated kaliophilite mine and sodium carbonate (Na2CO3) be uniformly mixed, then Calcination process is carried out using mechanical activation mode, maturing temperature is 850 DEG C, and the soaking time of roasting is 90min, obtains the first roasting Burn material.
Specifically, the dosage of sodium carbonate is SiO in kaliophilite mine2Just Na required for reacting completely2CO3The 1.2 of theoretical amount Times.Mechanical activation mode carries out roasting and refers to a kind of high-temperature machinery power conjunction of ball-milling treatment progress synchronous with high-temperature roasting reaction At the process of reaction, ball milling progress synchronous with two techniques of roasting can be realized, roasting is more efficient, can not only simplify work Skill, while maturing temperature and calcining time can also be reduced, energy consumption is substantially reduced, keeps reaction more abundant.It specifically can be using public affairs The number of opening is that high temperature energy ball grinder disclosed in the patent of CA101003024A is realized.
Part silicon oxygen bond and quartzy state silica and Na during roasting herein, in kaliophilite mine2CO3Occur anti- It answers, iron and Na in kaliophilite mine2CO3It does not react, the reaction equation being related to is mainly as follows:
KAlSi3O8+2Na2CO3→KAlSiO4+2Na2SiO3+2CO2↑
SiO2+Na2CO3=Na2SiO3+CO2↑
Wherein, KAlSiO is mainly contained in the first roasting material obtained4And Na2SiO3, can also contain a small amount of iron.? The tail gas generated during this calcination process is mainly CO2, absorption tail gas is carried out using ammonia solution in the present embodiment, obtains carbon It can be used in heavy aluminium in subsequent step S7 after sour hydrogen ammonium.Reaction equation involved in vent gas treatment process is as follows:
CO2+NH4OH=NH4HCO3
S3, dissolution filtering: the first roasting material is dissolved out with the sodium hydroxide solution that concentration is 0.2mol/L, leaching temperature It is 85 DEG C, dissolution time 90min.Then it is filtered separation, obtain the mixed solution of sodium metasilicate and sodium hydroxide and is contained The pre-desiliconizing slag of potassium and aluminium.
Wherein, the solid-liquid mass ratio of the first roasting material and NaOH solution is 1:4.Due to being roasted in step s 2 When, if Na2CO3Amount it is inadequate, KAlSi3O8It is also possible to there can be a small amount of amorphous SiO in part after decomposition2.Therefore, herein It is dissolved out on the one hand with NaOH solution primarily to first is roasted the Na in material in step S32SiO3It dissolves in the solution, separately It is on the one hand to promote a small amount of amorphous SiO2Further reaction generates Na2SiO3Into in solution, and then further increase Pre-desiliconizing rate.Simultaneously in this process in leaching, due to KAlSiO4It is more stable in itself hardly to react with NaOH solution, The a small amount of iron contained in first roasting material is hardly reacted with aqueous slkali.Therefore, the ingredient of pre-desiliconizing slag is mainly KAlSiO4, can also contain a small amount of iron.
S4, sulphur ammonium Desilication by roasting: by pre-desiliconizing slag and ammonium sulfate ((NH4)2SO4) be uniformly mixed, it is then roasted, is roasted Burning temperature is 475 DEG C, calcining time 75min, obtains the second roasting material.
Specifically, (NH4)2SO4Dosage be that ammonium sulfate required for potassium and aluminium in pre-desiliconizing slag reacts completely just is theoretical 1.2 times of amount.In this roasting process, the reaction equation being related to is mainly as follows:
2KAlSiO4+5(NH4)2SO4→K2SO4+2NH4Al(SO4)2+2H4SiO4+8NH3↑
Wherein, a small amount of iron contained in pre-desiliconizing slag also can some reacted with ammonium sulfate generate ferric sulfate, H4SiO4Again It can decompose and generate SiO2, therefore, K is mainly contained in the second obtained roasting material2SO4、NH4Al(SO4)2And SiO2, can also contain There are a small amount of ferric sulfate and a small amount of unreacted iron.Roasting obtained flue gas is mainly NH3, used in the present embodiment and use dilute sulfuric acid Absorb, obtained ammonium sulfate again after inspissation crystallizes return step S4 as raw material.It is anti-involved in fume treatment process Answer formula as follows:
2NH3+H2SO4→(NH4)2SO4
It is roasted further, in this embodiment carrying out second using ammonium sulfate again after roasting for the first time using sodium carbonate, Compared with existing middle use Ore Leaching, more efficient, the calcining time reduction of ammonium sulfate roasting.Conversion ratio is higher simultaneously, significantly Improve KAlSiO4Conversion ratio, so that its structure is sufficiently destroyed, and then potassium, aluminium and the silicon in kaliophilite mine are sufficiently divided From substantially increasing the utilization rate to potassium, aluminium and silicon in kaliophilite mine, realize high value added utilization.Sulfuric acid is used simultaneously After flue gas is obtained after ammonium roasting using dilute sulfuric acid absorption, obtained ammonium sulfate can also be recycled, and further improve utilization Rate.
S5, dissolution filtering: by second roasting material with it is water-soluble go out, leaching temperature be 80 DEG C, dissolution time 60min.So After be separated by filtration, obtain silicon slag and the aluminum ammonium sulfate solution containing potassium；
Wherein, the solid-liquid mass ratio of the second roasting material and water is 1:4, in this process in leaching primarily to making K2SO4、 NH4Al(SO4)2It is soluble in water with a small amount of ferric sulfate, and SiO2It is insoluble in water with a small amount of unreacted iron, therefore, this step It can be by K2SO4、NH4Al(SO4)2With a small amount of ferric sulfate and SiO2It is separated with a small amount of unreacted iron.What is obtained contains potassium Aluminum ammonium sulfate solution mainly include K2SO4And NH4Al(SO4)2, can also contain a small amount of ferric sulfate, mainly include in silicon slag SiO2With a small amount of unreacted iron.
S6, heavy aluminium: ammonium hydrogen carbonate is added in the aluminum ammonium sulfate solution containing potassium obtained to step S5, adjusts the pH value of solution To 5.6 heavy aluminium, aluminium is at this time with Al (OH)3Form precipitates completely.Then be filtered separation, obtain thick aluminum hydroxide precipitation and The mixed solution of potassium sulfate and ammonium sulfate.
Specifically, due to when pH is greater than 5.4 aluminium can precipitate completely, and potassium will not precipitate, therefore adjust in the present embodiment PH to 5.6 can aluminium and potassium separated.Key reaction formula involved in heavy aluminium is as follows:
2NH4Al(SO4)2+6NH4HCO3=4 (NH4)2SO4+2Al(OH)3↓+6CO2↑
Wherein, due to pH be greater than 3.6 when iron also can all precipitate, pH be 5.6 when iron precipitating completely, obtain It mainly include Al (OH) in thick aluminum hydroxide precipitation3, can also contain a small amount of iron hydroxide.The gas that the aluminium that sinks herein generates in the process Body is mainly CO2, absorbed, then crystallize to obtain crystals of sodium carbonate through inspissation, can be returned using NaOH solution in the present embodiment Step S2 is as raw material.Reaction equation involved in gas treating process is as follows:
CO2+ 2NaOH=Na2CO3+H2O
S7, fractional crystallization: the mixed solution of potassium sulfate and ammonium sulfate that step S6 is obtained carries out substep after circulation collection Crystallize preparing potassium sulfate crystal and ammonia sulfate crystal, obtained ammonia sulfate crystal can return step S4 be used as raw material, crystallization mother Liquid can be recycled.Optionally, by the mixed solution of the obtained potassium sulfate of step S6 and ammonium sulfate directly as composite chemical fertilizer It uses.
The potassium in kaliophilite mine is extracted at this time as a result, and potassium product has been prepared.
Contain ammonium salt solution further, in this embodiment selecting when heavy aluminium, with the heavy aluminium of existing middle selection sodium carbonate, the present embodiment In added value it is higher because using sink containing ammonium salt solution the ammonium sulfate that obtains after aluminium can with potassium sulfate solution together as Composite chemical fertilizer uses, and ammonia sulfate crystal is obtained after can also being crystallized again and can be recycled, utilization rate is substantially increased.
S8, alkali soluble: by thick aluminum hydroxide precipitation with concentration be 1mol/L sodium hydroxide solution 80 DEG C dissolve out, then into Row is separated by filtration, and obtains sodium aluminate solution and tailings.
Wherein, since a small amount of iron hydroxide contained in thick aluminum hydroxide precipitation can't be dissolved in aqueous slkali, To tailings be mainly a small amount of scum, be recycled after being collected.Therefore, the purpose of this step alkali soluble is mainly In order to which the iron tramp in thick aluminum hydroxide precipitation is separated with aluminium.Reaction equation involved in process in leaching is mainly as follows:
2Al(OH)3+2NaOH→2NaAlO2+4H2O
S9, carbon point: sodium aluminate solution is subjected to carbon point, separation is then filtered, obtains high-purity aluminium hydroxide and carbon Acid sodium solution.
Specifically, on the one hand carbon point i.e. carbonation decomposition can not introduce other impurities using carbon point, raising finally obtains Aluminium hydroxide purity, the CO in another aspect raw material2Raw material can be more saved by obtaining in other steps.If step If NaOH solution excess in S8, a small amount of NaOH solution is also had in obtained sodium aluminate solution, the reaction related generally to Formula is as follows:
2NaAlO2+CO2+3H2O→2Al(OH)3↓+Na2CO3
CO2+ 2NaOH=Na2CO3+H2O
Wherein, the sodium carbonate liquor obtained again after inspissation crystallizes can return step S2 as raw material or be used for step Causticization in S13.
The aluminium in kaliophilite mine is extracted at this time as a result, and aluminum hydroxide product has been prepared.
S10, alkali leaching: the silicon slag that step S5 is obtained is leached with the sodium hydroxide solution that concentration is 12.5mol/L, leaches temperature Degree is 135 DEG C, extraction time 40min.Leaching slurry is diluted with water again, makes Na in solution2SiO3Concentration dilution be 2mol/L.Then it is filtered separation, obtains sodium silicate solution and filter residue.
Wherein, the solid-liquid mass ratio of silicon slag and NaOH solution is 1:2.5.Due to using concentration for the hydrogen-oxygen of 12.5mol/L When change sodium solution is leached, the concentration of solution is higher, and on the one hand alkalinity is relatively strong, is difficult to be filtered, on the other hand molten The viscosity of liquid is larger, will also result in more difficult to filter.It to be first diluted after being leached therefore, it is necessary to silicon slag.Due in step s 4, Contained iron reacts not fully with ammonium sulfate, therefore can also contain a small amount of iron in obtained silicon slag, so in step slo Obtained filter residue is mainly scum.Key reaction formula involved in leaching process is as follows:
SiO2+ 2NaOH=Na2SiO3+H2O
S11, first time carbon point: the sodium silicate solution that step S10 is obtained carries out first time carbon point, and first time carbon divides terminal PH value be 11.5, be then filtered separation, obtain the mixed liquor and miscellaneous slag of sodium metasilicate and sodium carbonate.
Since in step slo, a small amount of iron contained in silicon slag is possible to have few a part into sodium hydroxide It in solution, while may also contain the substances such as calcium, manganese in kaliophilite mine, therefore may be used also in the obtained sodium silicate solution of step S10 The purity of contained sodium metasilicate in sodium silicate solution can be will affect containing impurity such as iron, calcium, manganese.Therefore, it first carries out in this step Carbon point, primarily to the above-mentioned impurity that may contain in removal sodium silicate solution.
Specifically, sodium metasilicate can carry out reacting generation sodium carbonate and silica with carbon dioxide and water during carbon point, And the SiO generated2, can be by impurity absorptions such as iron present in solution, calcium, manganese with very strong adsorptivity, and then realize and remove Miscellaneous effect.Final filtration impurity main component obtained after separation is SiO2And the impurity such as a small amount of iron, calcium, manganese, it obtains Miscellaneous slag can be to be mixed together in return step S10 with silicon slag.
S12, second of carbon point: carrying out second of carbon point for the mixed liquor of the obtained sodium metasilicate of step S11 and sodium carbonate, the Secondary carbon divides the pH value of terminal to be 9, is then filtered separation, obtains sodium carbonate liquor and white carbon black.
Specifically, in this step primarily to so that the sodium metasilicate in mixed liquor is sufficiently carried out carbon mitogenetic at micro-nano SiO2, i.e. white carbon black.And the sodium carbonate liquor obtained after being separated by filtration can the conduct in return step S2 after inspissation crystallizes again Raw material uses or for causticization in step S13.
The silicon in kaliophilite mine is extracted at this time as a result, and white carbon black product has been prepared.
Further, in this embodiment white carbon black product can be obtained using first time carbon point and second of carbon point, significantly Silicon product quality is improved, and the sodium carbonate obtained after carbon point can improve utilization rate with recycling.
S13, causticization: active lime lotion is added into sodium carbonate liquor and carries out causticizing reaction, causticization temperature is 90 DEG C, severe The change time is 40min, and the concentration of active lime lotion is 160g/L, is then filtered separation, obtains calcium carbonate and hydroxide Sodium solution.The reaction equation being related to is mainly as follows:
Ca(OH)2+Na2CO3→CaCO3↓+2NaOH
Wherein, the molar ratio of active lime cream and sodium carbonate are as follows: Ca (OH)2:Na2CO3=1.1:1 is carried out used when causticization Sodium carbonate liquor can come from step S2, S9 or S12.It may be containing not in the sodium hydroxide solution obtained after being separated by filtration The active lime cream of reaction, this amount of activated milk of lime can absorb CO in air2Precipitation of calcium carbonate is generated, then using filter After separation, available calcium carbonate and sodium hydroxide solution.
Ca(OH)2+CO2→CaCO3↓+H2O
Calcium carbonate product can also be obtained in the present embodiment as a result,.
Further, in this embodiment calcium carbonate product can be produced by causticization, and sodium carbonate used when causticization is molten Liquid can come from obtained in preceding step, and then substantially increase cyclic utilization rate, and the sodium hydroxide solution after causticization High value added utilization can also be realized with recycling.
S14, inspissation: the sodium hydroxide solution that step S13 is obtained carry out inspissation to the concentration of 12.5mol/L, obtains To sodium hydroxide inspissation solution return step S10 in leach silicon slag, the distilled water obtained after inspissation can recycle, significantly Cyclic utilization rate is improved, is economized on resources.
To sum up, kaliophilite mine is sufficiently decomposed by the way of double roasting in the present embodiment, first uses sodium carbonate First time roasting is carried out, and controls maturing temperature at 800~1000 DEG C, it can be by KAlSi in kaliophilite mine3O8Structure carry out Destruction makes its decomposition obtain KAlSiO4, realize pre-desiliconizing, and then the very stable kaliophilite mine of script is made to become active, it is more sharp It is reacted in subsequent with ammonium sulfate.Then second is carried out using ammonium sulfate again to roast, and control maturing temperature 350~ It 550 DEG C, can be by KAlSiO4Further reaction decomposes and obtains K2SO4、NH4Al(SO4)2And SiO2, realize and extract potassium aluminium, in turn Potassium, aluminium and silicon in kaliophilite mine is sufficiently separated.By that can be sufficiently destroyed the structure of kaliophilite mine after double roasting, make Potassium, aluminium and silicon therein are sufficiently separated.Therefore subsequently through can be obtained after being further processed high-purity aluminum hydroxide product, Potassium product, white carbon black product and calcium carbonate product substantially increase the utilization rate to potassium, aluminium and silicon in kaliophilite mine, Realize high value added utilization.
Specifically, then after heavy aluminium, alkali soluble and carbon point high-purity aluminum hydroxide product can will be obtained, then is tied step by step Available potassium product after crystalline substance is soaked by alkali and can obtain white carbon black product after carbon point twice, then can be with after causticization Obtain calcium carbonate product.And the by-product obtained during processing can be recycled recycling, has saved resource, has also mentioned The high utilization rate of resource.
In addition, using sodium carbonate roasting and ammonium sulfate roasting double roasting carry out decompose kaliophilite mine method, energy consumption compared with Low, the waste residue of generation is less, and the waste residue and tail gas that generate can recycle after processing, substantially increase the weight of resource It is multiple to utilize, meet the needs of economizing on resources and is environmental-friendly.Preparation method simple process simultaneously is suitble to large-scale production.
Embodiment 2
The present embodiment provides a kind of methods for comprehensively utilizing kaliophilite mine, specifically comprise the following steps:
S1, pretreatment: kaliophilite mine is pre-processed.Specifically, the chemistry of kaliophilite mine used in the present embodiment Composition specifically includes that SiO by weight percentage255.90%, Al2O319.83%, K2O 13.72%, Na20.02% He of O Fe2O37.65%.Preprocessing process is referring to embodiment 1, and details are not described herein.
S2, mixing and roasting pre-desiliconizing: by pretreated kaliophilite mine and Na2CO3It is uniformly mixed, then using machinery Activation method carries out calcination process, and maturing temperature is 875 DEG C, and the soaking time of roasting is 75min, obtains the first roasting material. Wherein, the dosage of sodium carbonate is SiO in kaliophilite mine2Just Na required for reacting completely2CO31.25 times of theoretical amount.To this roasting It burns the tail gas generated in treatment process to be collected, can be used for carbon point in step S9, step S11 and step S12.
S3, dissolution filtering: the first roasting material is dissolved out with the sodium hydroxide solution that concentration is 0.25mol/L, dissolution temperature Degree is 90 DEG C, dissolution time 60min.Then be filtered separation, obtain sodium metasilicate and sodium hydroxide mixed solution and Pre-desiliconizing slag containing potassium and aluminium.Wherein, the solid-liquid mass ratio of the first roasting material and NaOH solution is 1:5.
S4, sulphur ammonium Desilication by roasting: by pre-desiliconizing slag and (NH4)2SO4It is uniformly mixed, is then roasted, maturing temperature is 450 DEG C, calcining time 90min, obtain the second roasting material.Wherein, (NH4)2SO4Dosage be pre-desiliconizing slag in potassium and 1.25 times of ammonium sulfate theoretical amount required for aluminium reacts completely just.
S5, dissolution filtering: by second roasting material with it is water-soluble go out, leaching temperature be 85 DEG C, dissolution time 60min.So After be separated by filtration, obtain silicon slag and the aluminum ammonium sulfate solution containing potassium.Wherein, the solid-liquid mass ratio of the second roasting material and water For 1:4.
S6, heavy aluminium: being added ammonium hydroxide in the aluminum ammonium sulfate solution containing potassium obtained to step S5, adjust the pH value of solution to 5.6 carry out heavy aluminium, and aluminium is at this time with Al (OH)3Form precipitates completely.Then be filtered separation, obtain thick aluminum hydroxide precipitation with And the mixed solution of potassium sulfate and ammonium sulfate.Key reaction formula involved in heavy aluminium is as follows:
2NH4Al(SO4)2+6NH4OH=4 (NH4)2SO4+2Al(OH)3↓
S7, fractional crystallization: the mixed solution of potassium sulfate and ammonium sulfate that step S6 is obtained carries out substep after circulation collection Crystallize preparing potassium sulfate crystal and ammonia sulfate crystal, obtained ammonia sulfate crystal can return step S4 be used as raw material, crystallization mother Liquid can be recycled.The potassium in kaliophilite mine is extracted at this time as a result, and potassium product has been prepared.
S8, alkali soluble: thick aluminum hydroxide precipitation is dissolved out with the sodium hydroxide solution that concentration is 1.5mol/L at 90 DEG C, then It is filtered separation, obtains sodium aluminate solution and tailings.
S9, carbon point: sodium aluminate solution is subjected to carbon point, separation is then filtered, obtains high-purity aluminium hydroxide and carbon Acid sodium solution.The aluminium in kaliophilite mine is extracted at this time as a result, and aluminum hydroxide product has been prepared.
S10, alkali leaching: the silicon slag that step S5 is obtained is leached with the sodium hydroxide solution that concentration is 12mol/L, extraction temperature It is 125 DEG C, extraction time 45min.Leaching slurry is diluted with water again, makes Na in solution2SiO3Concentration dilution be 2mol/L.Then it is filtered separation, obtains sodium silicate solution and filter residue.Wherein, the solid-liquid mass ratio of silicon slag and NaOH solution For 1:2.5.
S11, first time carbon point: the sodium silicate solution that step S10 is obtained carries out first time carbon point, and first time carbon divides terminal PH value be 11, be then filtered separation, obtain the mixed liquor and miscellaneous slag of sodium metasilicate and sodium carbonate.
S12, second of carbon point: carrying out second of carbon point for the mixed liquor of the obtained sodium metasilicate of step S11 and sodium carbonate, the Secondary carbon divides the pH value of terminal to be 9, is then filtered separation, obtains sodium carbonate liquor and white carbon black.It has incited somebody to action at this time as a result, Silicon in kaliophilite mine extracts and white carbon black product has been prepared.
S13, causticization: active lime lotion is added into sodium carbonate liquor and carries out causticizing reaction, causticization temperature is 90 DEG C, severe The change time is 45min, and the concentration of active lime lotion is 180g/L, is then filtered separation, obtains calcium carbonate and hydroxide Sodium solution.
Wherein, the molar ratio of active lime cream and sodium carbonate is 1:1, can in the sodium hydroxide solution obtained after being separated by filtration CO can can be absorbed in air containing unreacted active lime cream, this amount of activated milk of lime2Precipitation of calcium carbonate is generated, so After being separated afterwards using filter, available calcium carbonate and sodium hydroxide solution.Calcium carbonate can also be obtained in the present embodiment as a result, Product.
S14, inspissation: the sodium hydroxide solution that step S13 is obtained carry out inspissation to the concentration of 12mol/L, obtains Sodium hydroxide inspissation solution return step S10 in leach silicon slag, the distilled water obtained after inspissation can recycle.
Embodiment 3
The present embodiment provides a kind of methods for comprehensively utilizing kaliophilite mine, specifically comprise the following steps:
S1, pretreatment: kaliophilite mine is pre-processed.Specifically, the chemistry of kaliophilite mine used in the present embodiment Composition specifically includes that SiO by weight percentage252.82%, Al2O322.50%, K2O 14.22%, Na20.54% He of O Fe2O34.88%.Preprocessing process is referring to embodiment 1, and details are not described herein.
S2, mixing and roasting pre-desiliconizing: by pretreated kaliophilite mine and Na2CO3It is uniformly mixed, then using machinery Activation method carries out calcination process, and maturing temperature is 800 DEG C, and the soaking time of roasting is 120min, obtains the first roasting material. Wherein, the dosage of sodium carbonate is SiO in kaliophilite mine2Just Na required for reacting completely2CO30.8 times of theoretical amount.To this roasting It burns the tail gas generated in treatment process to absorb using NaOH solution, obtains sodium carbonate liquor, then obtain sodium carbonate through evaporative crystallization Return step S2 is as raw material after crystal.Reaction equation involved in vent gas treatment process is as follows:
CO2+ 2NaOH=Na2CO3+H2O
S3, dissolution filtering: the first roasting material is dissolved out with the sodium hydroxide solution that concentration is 0.4mol/L, leaching temperature It is 50 DEG C, dissolution time 120min.Then it is filtered separation, obtain the mixed solution of sodium metasilicate and sodium hydroxide and is contained The pre-desiliconizing slag of potassium and aluminium.Wherein, the solid-liquid mass ratio of the first roasting material and NaOH solution is 1:2.
S4, sulphur ammonium Desilication by roasting: by pre-desiliconizing slag and (NH4)2SO4It is uniformly mixed, is then roasted, maturing temperature is 450 DEG C, calcining time 90min, obtain the second roasting material.Wherein, (NH4)2SO4Dosage be pre-desiliconizing slag in potassium and 1.0 times of ammonium sulfate theoretical amount required for aluminium reacts completely just.
S5, dissolution filtering: by second roasting material with it is water-soluble go out, leaching temperature be 60 DEG C, dissolution time 100min.So After be separated by filtration, obtain silicon slag and the aluminum ammonium sulfate solution containing potassium.Wherein, the solid-liquid mass ratio of the second roasting material and water For 1:5.
S6, heavy aluminium: being added ammonium hydroxide in the aluminum ammonium sulfate solution containing potassium obtained to step S5, adjust the pH value of solution to 5.4 carry out heavy aluminium, and aluminium is at this time with Al (OH)3Form precipitates completely.Then be filtered separation, obtain thick aluminum hydroxide precipitation with And the mixed solution of potassium sulfate and ammonium sulfate.
S7, fractional crystallization: the mixed solution of potassium sulfate and ammonium sulfate that step S6 is obtained carries out substep after circulation collection Crystallize preparing potassium sulfate crystal and ammonia sulfate crystal, obtained ammonia sulfate crystal can return step S4 be used as raw material, crystallization mother Liquid can be recycled.The potassium in kaliophilite mine is extracted at this time as a result, and potassium product has been prepared.
S8, alkali soluble: thick aluminum hydroxide precipitation is dissolved out with the sodium hydroxide solution that concentration is 0.5mol/L at 70 DEG C, then It is filtered separation, obtains sodium aluminate solution and tailings.
S9, carbon point: sodium aluminate solution is subjected to carbon point, separation is then filtered, obtains high-purity aluminium hydroxide and carbon Acid sodium solution.The aluminium in kaliophilite mine is extracted at this time as a result, and aluminum hydroxide product has been prepared.
S10, alkali leaching: the silicon slag that step S5 is obtained is leached with the sodium hydroxide solution that concentration is 4mol/L, extraction temperature It is 75 DEG C, extraction time 80min.Leaching slurry is diluted with water again, makes Na in solution2SiO3Concentration dilution be 1mol/L.Then it is filtered separation, obtains sodium silicate solution and filter residue.Wherein, the solid-liquid mass ratio of silicon slag and NaOH solution For 1:6.
S11, first time carbon point: the sodium silicate solution that step S10 is obtained carries out first time carbon point, and first time carbon divides terminal PH value be 11.2, be then filtered separation, obtain the mixed liquor and miscellaneous slag of sodium metasilicate and sodium carbonate.
S12, second of carbon point: carrying out second of carbon point for the mixed liquor of the obtained sodium metasilicate of step S11 and sodium carbonate, the Secondary carbon divides the pH value of terminal to be 9, is then filtered separation, obtains sodium carbonate liquor and white carbon black.It has incited somebody to action at this time as a result, Silicon in kaliophilite mine extracts and white carbon black product has been prepared.
S13, causticization: active lime lotion is added into sodium carbonate liquor and carries out causticizing reaction, causticization temperature is 80 DEG C, severe The change time is 60min, and the concentration of active lime lotion is 140g/L, is then filtered separation, obtains calcium carbonate and hydroxide Sodium solution.
Wherein, the molar ratio of active lime cream and sodium carbonate is 0.9:1, in the sodium hydroxide solution obtained after being separated by filtration CO may can be absorbed in air containing unreacted active lime cream, this amount of activated milk of lime2Precipitation of calcium carbonate is generated, Then after being separated using filter, available calcium carbonate and sodium hydroxide solution.Carbonic acid can also be obtained in the present embodiment as a result, Calcium product.
S14, inspissation: the sodium hydroxide solution that step S13 is obtained carry out inspissation to the concentration of 4mol/L, obtains Silicon slag is leached in sodium hydroxide inspissation solution return step S10, the distilled water obtained after inspissation can recycle.
Embodiment 4
The present embodiment provides a kind of methods for comprehensively utilizing kaliophilite mine, specifically comprise the following steps:
S1, pretreatment: kaliophilite mine is pre-processed.Specifically, the chemistry of kaliophilite mine used in the present embodiment Composition specifically includes that SiO by weight percentage260.56%, Al2O323.42%, K2O 11.98%, Na20.25% He of O Fe2O32.68%.Preprocessing process is referring to embodiment 1, and details are not described herein.
S2, mixing and roasting pre-desiliconizing: by pretreated kaliophilite mine and Na2CO3It is uniformly mixed, then using machinery Activation method carries out calcination process, and maturing temperature is 900 DEG C, and the soaking time of roasting is 70min, obtains the first roasting material. Wherein, the dosage of sodium carbonate is SiO in kaliophilite mine2Just Na required for reacting completely2CO31.0 times of theoretical amount.To this roasting The tail gas generated in treatment process is burnt to absorb using ammonia solution.
S3, dissolution filtering: by first roasting material with it is water-soluble go out, leaching temperature be 70 DEG C, dissolution time 70min.So After be separated by filtration, obtain sodium silicate solution and the pre-desiliconizing slag containing potassium and aluminium.Wherein, the first roasting material and water are consolidated Liquid mass ratio is 1:3.
S4, sulphur ammonium Desilication by roasting: by pre-desiliconizing slag and (NH4)2SO4It is uniformly mixed, is then roasted, maturing temperature is 500 DEG C, calcining time 60min, obtain the second roasting material.Wherein, (NH4)2SO4Dosage be pre-desiliconizing slag in potassium and 1.4 times of ammonium sulfate theoretical amount required for aluminium reacts completely just.
S5, dissolution filtering: by second roasting material with it is water-soluble go out, leaching temperature be 70 DEG C, dissolution time 90min.So After be separated by filtration, obtain silicon slag and the aluminum ammonium sulfate solution containing potassium.Wherein, the solid-liquid mass ratio of the second roasting material and water For 1:3.
S6, heavy aluminium: ammonium hydrogen carbonate is added in the aluminum ammonium sulfate solution containing potassium obtained to step S5, adjusts the pH value of solution Heavy aluminium is carried out to 5.5, aluminium is at this time with Al (OH)3Form precipitates completely.Then it is filtered separation, obtains thick aluminum hydroxide precipitation And the mixed solution of potassium sulfate and ammonium sulfate.
S7, fractional crystallization: the mixed solution of potassium sulfate and ammonium sulfate that step S6 is obtained carries out substep after circulation collection Crystallize preparing potassium sulfate crystal and ammonia sulfate crystal, obtained ammonia sulfate crystal can return step S4 be used as raw material, crystallization mother Liquid can be recycled.The potassium in kaliophilite mine is extracted at this time as a result, and potassium product has been prepared.
S8, alkali soluble: thick aluminum hydroxide precipitation is dissolved out with the sodium hydroxide solution that concentration is 1.5mol/L at 95 DEG C, then It is filtered separation, obtains sodium aluminate solution and tailings.
S9, carbon point: sodium aluminate solution is subjected to carbon point, separation is then filtered, obtains high-purity aluminium hydroxide and carbon Acid sodium solution.The aluminium in kaliophilite mine is extracted at this time as a result, and aluminum hydroxide product has been prepared.
S10, alkali leaching: the silicon slag that step S5 is obtained is leached with the sodium hydroxide solution that concentration is 8mol/L, extraction temperature It is 100 DEG C, extraction time 60min.Leaching slurry is diluted with water again, makes Na in solution2SiO3Concentration dilution be 2mol/L.Then it is filtered separation, obtains sodium silicate solution and filter residue.Wherein, the solid-liquid mass ratio of silicon slag and NaOH solution For 1:3.
S11, first time carbon point: the sodium silicate solution that step S10 is obtained carries out first time carbon point, and first time carbon divides terminal PH value be 11.5, be then filtered separation, obtain the mixed liquor and miscellaneous slag of sodium metasilicate and sodium carbonate.
S12, second of carbon point: carrying out second of carbon point for the mixed liquor of the obtained sodium metasilicate of step S11 and sodium carbonate, the Secondary carbon divides the pH value of terminal to be 8.5, is then filtered separation, obtains sodium carbonate liquor and white carbon black.As a result, at this time Silicon in kaliophilite mine is extracted and white carbon black product has been prepared.
S13, causticization: active lime lotion is added into sodium carbonate liquor and carries out causticizing reaction, causticization temperature is 95 DEG C, severe The change time is 30min, and the concentration of active lime lotion is 200g/L, is then filtered separation, obtains calcium carbonate and hydroxide Sodium solution.
Wherein, the molar ratio of active lime cream and sodium carbonate is 1:1, can in the sodium hydroxide solution obtained after being separated by filtration CO can can be absorbed in air containing unreacted active lime cream, this amount of activated milk of lime2Precipitation of calcium carbonate is generated, so After being separated afterwards using filter, available calcium carbonate and sodium hydroxide solution.Calcium carbonate can also be obtained in the present embodiment as a result, Product.
S14, inspissation: the sodium hydroxide solution that step S13 is obtained carry out inspissation to the concentration of 8mol/L, obtains Silicon slag is leached in sodium hydroxide inspissation solution return step S10.The distilled water obtained after inspissation can recycle.
Embodiment 5
The present embodiment provides a kind of methods for comprehensively utilizing kaliophilite mine, specifically comprise the following steps:
S1, pretreatment: kaliophilite mine is pre-processed.Specifically, the chemistry of kaliophilite mine used in the present embodiment Composition specifically includes that SiO by weight percentage266.24%, Al2O321.02%, K2O 12.65%, Na20.15% He of O Fe2O35.24%.Preprocessing process is referring to embodiment 1, and details are not described herein.
S2, mixing and roasting pre-desiliconizing: by pretreated kaliophilite mine and Na2CO3It is uniformly mixed, then using machinery Activation method carries out calcination process, and maturing temperature is 850 DEG C, and the soaking time of roasting is 90min, obtains the first roasting material. Wherein, the dosage of sodium carbonate is SiO in kaliophilite mine2Just Na required for reacting completely2CO31.4 times of theoretical amount.To this roasting The tail gas generated in treatment process is burnt to absorb using NaOH solution.
S3, dissolution filtering: by first roasting material with it is water-soluble go out, leaching temperature be 95 DEG C, dissolution time 60min.So After be separated by filtration, obtain sodium silicate solution and the pre-desiliconizing slag containing potassium and aluminium.Wherein, the first roasting material and water are consolidated Liquid mass ratio is 1:6.
S4, sulphur ammonium Desilication by roasting: by pre-desiliconizing slag and (NH4)2SO4It is uniformly mixed, is then roasted, maturing temperature is 475 DEG C, calcining time 75min, obtain the second roasting material.Wherein, (NH4)2SO4Dosage be pre-desiliconizing slag in potassium and 1.6 times of ammonium sulfate theoretical amount required for aluminium reacts completely just.
S5, dissolution filtering: by second roasting material with it is water-soluble go out, leaching temperature be 90 DEG C, dissolution time 40min.So After be separated by filtration, obtain silicon slag and the aluminum ammonium sulfate solution containing potassium.Wherein, the solid-liquid mass ratio of the second roasting material and water For 1:6.
S6, heavy aluminium: ammonium hydrogen carbonate is added in the aluminum ammonium sulfate solution containing potassium obtained to step S5, adjusts the pH value of solution Heavy aluminium is carried out to 5.6, aluminium is at this time with Al (OH)3Form precipitates completely.Then it is filtered separation, obtains thick aluminum hydroxide precipitation And the mixed solution of potassium sulfate and ammonium sulfate.
S7, fractional crystallization: the mixed solution of potassium sulfate and ammonium sulfate that step S6 is obtained carries out substep after circulation collection Crystallize preparing potassium sulfate crystal and ammonia sulfate crystal, obtained ammonia sulfate crystal can return step S4 be used as raw material, crystallization mother Liquid can be recycled.The potassium in kaliophilite mine is extracted at this time as a result, and potassium product has been prepared.
S8, alkali soluble: by thick aluminum hydroxide precipitation with concentration be 2mol/L sodium hydroxide solution 95 DEG C dissolve out, then into Row is separated by filtration, and obtains sodium aluminate solution and tailings.
S9, carbon point: sodium aluminate solution is subjected to carbon point, separation is then filtered, obtains high-purity aluminium hydroxide and carbon Acid sodium solution.The aluminium in kaliophilite mine is extracted at this time as a result, and aluminum hydroxide product has been prepared.
S10, alkali leaching: the silicon slag that step S5 is obtained is leached with the sodium hydroxide solution that concentration is 15mol/L, extraction temperature It is 145 DEG C, extraction time 30min.Leaching slurry is diluted with water again, makes Na in solution2SiO3Concentration dilution be 2.5mol/L.Then it is filtered separation, obtains sodium silicate solution and filter residue.Wherein, the solid-liquid quality of silicon slag and NaOH solution Than for 1:2.2.
S11, first time carbon point: the sodium silicate solution that step S10 is obtained carries out first time carbon point, and first time carbon divides terminal PH value be 11.5, be then filtered separation, obtain the mixed liquor and miscellaneous slag of sodium metasilicate and sodium carbonate.
S12, second of carbon point: carrying out second of carbon point for the mixed liquor of the obtained sodium metasilicate of step S11 and sodium carbonate, the Secondary carbon divides the pH value of terminal to be 9, is then filtered separation, obtains sodium carbonate liquor and white carbon black.It has incited somebody to action at this time as a result, Silicon in kaliophilite mine extracts and white carbon black product has been prepared.
S13, causticization: active lime lotion is added into sodium carbonate liquor and carries out causticizing reaction, causticization temperature is 80 DEG C, severe The change time is 60min, and the concentration of active lime lotion is 240g/L, is then filtered separation, obtains calcium carbonate and hydroxide Sodium solution.
Wherein, the molar ratio of active lime cream and sodium carbonate is 1.2:1, in the sodium hydroxide solution obtained after being separated by filtration CO may can be absorbed in air containing unreacted active lime cream, this amount of activated milk of lime2Precipitation of calcium carbonate is generated, Then after being separated using filter, available calcium carbonate and sodium hydroxide solution.Carbonic acid can also be obtained in the present embodiment as a result, Calcium product.
S14, inspissation: the sodium hydroxide solution that step S13 is obtained carry out inspissation to the concentration of 15mol/L, obtains Sodium hydroxide inspissation solution return step S10 in leach silicon slag.The distilled water obtained after inspissation can recycle.
More than, it is only presently preferred embodiments of the present invention, is not the limitation for doing other forms to invention, it is any to be familiar with Professional and technical personnel is changed or is modified as the equivalence enforcement of equivalent variations possibly also with the technology contents of the disclosure above Example.But without departing from the technical solutions of the present invention, according to the technical essence of the invention to the above embodiments What simple modification, equivalent variations and remodeling, still falls within the protection scope of technical solution of the present invention.

Claims (10)

1. a kind of method for comprehensively utilizing kaliophilite mine, which comprises the steps of:
S1, kaliophilite mine is pre-processed；
S2, pretreated kaliophilite mine is uniformly mixed with sodium carbonate, then carries out calcination process, obtains the first calcining matter Material, maturing temperature are 800~1000 DEG C, and the soaking time of roasting is 50~150min；
S3, by it is described first roasting material water or sodium hydroxide solution dissolve out, be then filtered separation, obtain containing potassium with The pre-desiliconizing slag of aluminium；
S4, the pre-desiliconizing slag is uniformly mixed with ammonium sulfate, is then roasted, obtains the second roasting material, maturing temperature It is 350~550 DEG C, calcining time is 50~150min；
S5, by second roasting material with it is water-soluble out, be then filtered separation, obtain silicon slag and the aluminum ammonium sulfate solution containing potassium；
It being added in S6, the aluminum ammonium sulfate solution containing potassium obtained to step S5 and contains ammonium salt solution, the pH value for adjusting solution carries out heavy aluminium, Then it is filtered separation, obtains the mixed solution of thick aluminum hydroxide precipitation and potassium sulfate and ammonium sulfate；
The mixed solution of S7, the potassium sulfate that step S6 is obtained and ammonium sulfate carry out fractional crystallization preparing potassium sulfate after circulation collection Crystal and ammonia sulfate crystal, the ammonia sulfate crystal return step S4 is as raw material；Or the potassium sulfate for obtaining step S6 and The mixed solution of ammonium sulfate is used directly as composite chemical fertilizer.
2. the method for comprehensive utilization kaliophilite mine as described in claim 1, which is characterized in that
In step sl, the pretreatment includes being crushed the kaliophilite mine, then that the granularity of the kaliophilite mine is levigate To 80 μm or less；
In step s 2, the dosage of the sodium carbonate is that silica reacts required sodium carbonate reason completely just in kaliophilite mine 0.8~1.4 times of stoichiometric, is roasted using mechanical activation mode；
In step s3, when using sodium hydroxide solution to be dissolved out the concentration of sodium hydroxide solution used for 0.01~ 0.4mol/L, solid-liquid mass ratio when dissolution are 1:2~6, and leaching temperature is 50~95 DEG C, and dissolution time is 20~120min.
3. the method for comprehensive utilization kaliophilite mine as claimed in claim 2, which is characterized in that
In step s 4, the dosage of the ammonium sulfate is that potassium and aluminium in the pre-desiliconizing slag react required sulfuric acid completely just 1.0~1.6 times of ammonium theoretical amount；
In step s 5, the solid-liquid mass ratio of the second roasting material and water is 1:2~6, and leaching temperature is 60~90 DEG C, molten The time is 30~120min out；
In step s 6, the ammonium salt solution that contains is at least one of ammonium hydroxide and ammonium hydrogen carbonate, and the pH value is 5.4~5.6.
4. the method for comprehensive utilization kaliophilite mine as described in any one of claims 1-3, which is characterized in that after step S6 Further include following steps:
S8, the thick aluminum hydroxide precipitation is dissolved out with aqueous slkali, is then filtered separation, obtains sodium aluminate solution and tail Slag；
S9, the sodium aluminate solution is subjected to carbon point, is then filtered separation, obtains aluminium hydroxide and sodium carbonate liquor.
5. as claimed in claim 4 comprehensive utilization kaliophilite mine method, which is characterized in that further include after step s 5 as Lower step:
S10, the obtained silicon slag of step S5 is leached, then be diluted to slurry is leached with sodium hydroxide solution, was then carried out Filter separation, obtains sodium silicate solution and filter residue；
S11, the sodium silicate solution for obtaining step S10 carry out first time carbon point, are then filtered separation, obtain sodium metasilicate with The mixed liquor of sodium carbonate and miscellaneous slag；
S12, the mixed liquor of the obtained sodium metasilicate of step S11 and sodium carbonate is subjected to second of carbon point, is then filtered separation, Obtain sodium carbonate liquor and white carbon black.
6. the method for comprehensive utilization kaliophilite mine as claimed in claim 5, which is characterized in that
In step s 8, aqueous slkali is sodium hydroxide solution, and the concentration of sodium hydroxide solution is 0.5~2mol/L, leaching temperature It is 70~95 DEG C；
In step slo, the concentration of sodium hydroxide solution is 4~15mol/L, the solid-liquid matter of the silicon slag and sodium hydroxide solution It measures than being 1:1.5~5, extraction temperature is 75~145 DEG C, and extraction time is 30~90min, silicon in solution after leaching pulp dilution The concentration of sour sodium is 1~2.5mol/L；
In step s 11, first time carbon divides the pH value of terminal to be 11.0~11.5；
In step s 12, second of carbon divides the pH value of terminal to be 8.5~9.
7. the method for comprehensive utilization kaliophilite mine as claimed in claim 6, which is characterized in that
In step s 2, the tail gas that calcination process generates is absorbed using sodium hydroxide solution, and then evaporative crystallization obtains sodium carbonate Return step S2 is as raw material after crystal；Or the tail gas generated to calcination process is collected, and is used for step S9, step S11 With carbon in step S12 point；Or the tail gas that calcination process generates is absorbed using ammonia solution, is used for step S6 after obtaining ammonium hydrogen carbonate In sink aluminium；
In step s 4, the flue gas roasted is absorbed using dilute sulfuric acid, and obtained ammonium sulfate returns after inspissation crystallizes again Step S4 is returned as raw material；
In step s 6, it is absorbed using the gas generated during bicarbonate ammonia-sinking aluminium using sodium hydroxide solution, obtained carbon Acid sodium solution again after inspissation crystallizes return step S2 as raw material；
In step s 11, the miscellaneous slag obtained in return step S10 with silicon slag for mixing together.
8. the method for comprehensive utilization kaliophilite mine as claimed in claim 7, which is characterized in that further include after step s 12 Following steps:
S13, active lime lotion progress causticizing reaction is added into sodium carbonate liquor, is then filtered separation, obtains carbonic acid Calcium and sodium hydroxide solution；
S14, the sodium hydroxide solution for obtaining step S13 carry out inspissation to a certain concentration, obtained sodium hydroxide inspissation Silicon slag is leached in solution return step S10.
9. the method for comprehensive utilization kaliophilite mine as claimed in claim 8, which is characterized in that
In step s 13, the molar ratio of active lime cream and sodium carbonate is 0.9~1.2:1, and causticization temperature is 50~95 DEG C, severe The change time is 10~90min, and the concentration of active lime lotion is 140~240g/L.
10. the method for comprehensive utilization kaliophilite mine as claimed in claim 9, which is characterized in that
In step s 9, return step S2 as raw material or is used for step to the sodium carbonate liquor obtained after inspissation crystallizes again Causticization in S13；
In step s 12, the sodium carbonate liquor obtained is used or is used as raw material in return step S2 after inspissation crystallizes again The causticization in step S13；
In step S14, the distilled water obtained after inspissation is for recycling.
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Cited By (2)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN111100988A (en) * 2019-12-13 2020-05-05 潘爱芳 Method for reducing and utilizing multi-metal tailings
CN111620353A (en) * 2020-05-18 2020-09-04 山西宝盛远华新材料股份有限公司 Method for extracting potassium salt from potassium feldspar

Citations (3)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN1184091A (en) * 1996-11-29 1998-06-10 中国地质大学（北京） Potash fertilizer producing method by using potassium-rock
CN101993256A (en) * 2009-08-13 2011-03-30 中国地质大学（北京） Method for preparing agricultural potassium sulfate from potassium-rich rock
CN102557050A (en) * 2011-12-20 2012-07-11 昆明冶金研究院 New process for comprehensively utilizing potassium feldspar

Patent Citations (3)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN1184091A (en) * 1996-11-29 1998-06-10 中国地质大学（北京） Potash fertilizer producing method by using potassium-rock
CN101993256A (en) * 2009-08-13 2011-03-30 中国地质大学（北京） Method for preparing agricultural potassium sulfate from potassium-rich rock
CN102557050A (en) * 2011-12-20 2012-07-11 昆明冶金研究院 New process for comprehensively utilizing potassium feldspar

Non-Patent Citations (1)

* Cited by examiner, † Cited by third party
Title
刘佳囡 等: "钾长石综合利用工艺研究", 《矿产综合利用》 *

Cited By (2)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN111100988A (en) * 2019-12-13 2020-05-05 潘爱芳 Method for reducing and utilizing multi-metal tailings
CN111620353A (en) * 2020-05-18 2020-09-04 山西宝盛远华新材料股份有限公司 Method for extracting potassium salt from potassium feldspar

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