WO2015176429A1 - Method for extracting vanadium by leaching vanadium-containing raw material fired clinkers with ammonium bicarbonate solution - Google Patents
Method for extracting vanadium by leaching vanadium-containing raw material fired clinkers with ammonium bicarbonate solution Download PDFInfo
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- C01G31/00—Compounds of vanadium
Definitions
- the invention belongs to the technical field of vanadium chemical metallurgy, and in particular relates to a method for extracting vanadium from a clinker carbon ammonium solution after calcination of a vanadium-containing raw material.
- Vanadium-titanium magnetite, vanadium slag, stone coal and other vanadium-containing raw materials The traditional vanadium extraction method is vanadium extraction by roasting-water immersion/acid leaching method.
- the vanadium-containing raw material sodium roasting-water leaching is the mainstream method for extracting vanadium.
- the basic principle of the sodium roasting process is to convert the low-valent vanadium in the vanadium-containing raw material into water-soluble five by high-temperature calcination (750-850 ° C) with sodium salt such as Na 2 CO 3 , NaCl, Na 2 SO 4 as an additive.
- the sodium salt of vanadium is directly immersed in the sodium calcined product to obtain a vanadium-containing leach solution, and then an ammonium salt is added to prepare ammonium vanadate precipitation, and the vanadium oxide product is obtained by reduction roasting.
- the recovery rate of vanadium in the sodium roasting process is low.
- the recovery rate of vanadium in a single roasting is about 70%.
- the recovery rate of vanadium is only 80%, and it needs to be calcined several times, and the energy consumption is high.
- the subsequent ammonia precipitation process will result in high salinity ammonia-containing wastewater.
- the ammonia-containing wastewater not only pollutes the environment, but also has a high cost of treatment.
- calcification roasting-acid leaching vanadium has been widely concerned as a relatively clean vanadium extraction process.
- the method uses lime, limestone, dolomite and other calcareous raw materials instead of sodium salt additives to carry out high temperature roasting of vanadium slag, convert vanadium into water-insoluble calcium vanadate, and then dilute with dilute acid, and vanadium in the leachate is hydrolyzed by vanadium. Or ion exchange recovery. This method has been applied industrially in the Tula ferrous metallurgical plant in the Soviet Union.
- Pangang Group (patent CN 101161831A) has made progress in product impurity removal and purification, and completed the pilot test.
- Calcification roasting avoids chlorine/sulfur contamination of conventional sodium roasting, but vanadium conversion for calcification roasting is still less than 80% due to mass transfer barriers that are also limited by the calcination process.
- CN 101899582A proposes a method for extracting vanadium pentoxide from vanadium slag.
- the vanadium slag is calcined at 800-1000 ° C, and the clinker is subjected to alkali leaching of 20 to 50% alkali concentration, and the vanadic acid is obtained by silicon removal and cooling crystallization.
- the sodium crystals are then converted to vanadium by vanadium, which is converted to acid and ammonium salts.
- the method has the problem that impurities are difficult to remove and difficult to filter, and the problem of acid ammonia nitrogen wastewater can not be avoided, and the vanadium recovery rate is less than 90%.
- CN 1082617A proposes to treat the high-temperature vanadium slag obtained by blowing directly at 900-1300 ° C to promote the oxidation of low-valent vanadium in the slag into pentavalent vanadium. After the slag is cooled and crushed, at a certain temperature, alkali concentration, oxygen The vanadium in the slag is leached under partial pressure. The method does not need to be calcined at a high temperature after the vanadium slag is cooled down, the energy consumption is greatly reduced, and the environmental pollution caused by the sodium roasting is avoided, but the metal iron cannot be recovered, and the vanadium recovery rate is low. The problem.
- one of the objects of the present invention is to provide a method for leaching vanadium from a clinker-containing carbon ammonium solution after calcination of a vanadium-containing raw material.
- the vanadium leaching rate is high, the leaching temperature is low, the leaching conditions are mild, and other impurities in the vanadium-containing raw material do not enter the leachate, and subsequent ammonium vanadate separation is simple. Since the roasting process does not use sodium salt, the operation is easy to control, no harmful kiln gas is generated, and no salty wastewater is produced.
- a method for leaching vanadium by clinker carbon ammonium solution after calcination of a vanadium-containing raw material pulverizing the crust-containing raw material in a solution of vanadium in a solution of vanadium, and causing vanadium to enter the solution in the form of ammonium vanadate to obtain leaching slag and Vanadium solution.
- the vanadium-containing raw material is calcined, the trivalent vanadium is converted into a leachable pentavalent vanadium during the roasting process, and the pentavalent vanadium can be reacted with the ammonium bicarbonate solution to form ammonium vanadate, thereby extracting the vanadium from the raw material.
- the vanadium-containing raw material is vanadium-titanium magnetite or stone. a mixture of one or more of coal, vanadium slag or a vanadium-containing catalyst; wherein the vanadium-containing catalyst may be a catalyst before use or a catalyst that cannot be reused after use;
- the calcination is one or a combination of two or more of blank calcination (ie, no additive calcination) and calcification calcination.
- the ammonium bicarbonate solution refers to NH 4 + /HCO 3 - /OH - , NH 4 + /CO 3 2- /OH - or NH 4 + /HCO 3 - /CO 3 2- /OH - solution, which means that it can be NH 4 HCO 3 solution, (NH 4 ) 2 CO 3 solution, mixed solution of NH 4 HCO 3 and (NH 4 ) 2 CO 3 , ammonia water and NH 4 HCO A mixed solution of 3, a mixed solution of ammonia and (NH 4 ) 2 CO 3 , a mixture of ammonia and NH 4 HCO 3 and (NH 4 ) 2 CO 3 can also be obtained by carbonizing a CO 2 gas in an ammonia water. .
- the concentration of the ammonium carbonate solution is determined by the molar concentration of each ion in the solution: the concentration of NH 4 + is 0.5-6 mol/L, and the concentration of HCO 3 - is 0.5-6 mol/L.
- the concentration of CO 3 2- is 0.5-5 mol/L, which can be obtained by adding different solution combinations.
- the concentration of NH 4 + /HCO 3 - /CO 3 2- in the ammonium carbonate solution is less than 0.5mol/L, the extraction rate of vanadium is very low, and the concentration of NH 4 + /HCO 3 - is higher than 6mol/L, CO 3 2-
- the concentration of vanadium is more than 5mol/L, and the leaching rate of vanadium is basically unchanged during leaching. If the concentration of ammonium carbonate solution is increased, the cost of vanadium extraction will be greatly increased.
- the concentration of the ammonium carbonate solution in the present invention is: NH 4 + concentration is The concentration of 0.5 to 6 mol/L, HCO 3 - is 0.5 to 6 mol/L, and the concentration of CO 3 2- is 0.5 to 5 mol/L.
- the temperature of the leaching is from normal temperature to 150 ° C, for example, 25 ° C, 35 ° C, 50 ° C, 80 ° C, 95 ° C, 120 ° C, 150 ° C, etc., the temperature is too low.
- the vanadium leaching efficiency of the clinker-containing raw material after calcination is low, and the leaching rate of vanadium of various roasting clinker at 150 ° C is already high, and it is meaningless to raise the leaching temperature. Therefore, the leaching temperature of the present invention is selected from normal temperature to 150 °C.
- °C is preferably 50 to 95 ° C, and more preferably 60 to 75 ° C.
- the leaching time is 0.5 h or more, for example, 1 h, 2 h, 3 h, 4 h, etc., and the leaching time is too short, such as less than 0.5 h, and the extraction efficiency of vanadium is low, preferably. For 1 to 3 hours.
- the mass ratio of the liquid solid product of the ammonium carbonate solution to the clinker-containing raw material after calcination is 2:1 to 10:1 m 3 /t, for example, 4 : 1m 3 /t, 7:1m 3 /t, 9:1m 3 /t, etc.
- the liquid-solid ratio of the reaction liquid below 2m 3 /t is poor, and the liquid-solid ratio is 10m 3 /t.
- the amount of vanadium dissolved is sufficient, and it is no longer meaningful to increase the liquid-solid ratio, and is preferably from 3:1 to 6:1 m 3 /t.
- the method of the present invention comprises the steps of: leaching vanadium in a solution of vanadium containing raw materials in a solution of ammonium carbonate, the concentration of the ammonium carbonate solution is: NH 4 + concentration of 0.5-6 mol/L, HCO 3 - concentration 0.5 ⁇ 6mol / L, CO 3 2- concentration 0.5 ⁇ 5mol / L, leaching temperature is normal temperature ⁇ 150 ° C, leaching time is more than 0.5h, carbon ammonium solution and vanadium containing raw materials after roasting clinker The liquid-solid ratio is from 2:1 to 10:1 m 3 /t. After leaching, vanadium is introduced into the solution as ammonium vanadate to obtain a leaching residue and a vanadium-containing solution.
- the vanadium leaching rate is high, and impurities do not enter the vanadium-containing liquid, and the simple and clean separation of vanadium can be realized.
- the clinker obtained by blank calcination of vanadium slag with a V 2 O 5 content of 11.6% at 900 ° C was leached in a solution of NH 4 + 6 mol/L, HCO 3 - 6 mol/L ammonium nitrate solution, and the leaching temperature was obtained.
- the solid solution mass ratio of the leachate was 10 m 3 /t
- the leaching time was 3 h
- the leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue.
- the recovery of vanadium in the vanadium slag was 94.2%.
- the vanadium extract was extracted from the ammonium carbonate solution of /L, the leaching temperature was 75 ° C, the solid solution mass ratio of the leachate was 8 m 3 /t, and the leaching time was 2 h, and the leaching residue and the vanadium-containing solution were obtained.
- the leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue.
- the recovery of vanadium in the vanadium slag was 90.2%.
- the clinker obtained by calcination of vanadium slag with a V 2 O 5 content of 14.7% at 800 ° C is at a concentration of 5 mol/L of NH 4 + , a concentration of 2 mol/L of HCO 3 and a concentration of 1 mol/L of CO 3 2- .
- the vanadium extraction solution was leached with vanadium, the leaching temperature was 150 ° C, the solid matter mass ratio of the leaching solution was 5 m 3 /t, and the leaching time was 2.5 h, and the leaching residue and the ammonia leaching vanadium containing solution were obtained.
- the leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue.
- the recovery of vanadium in the vanadium slag was 93.4%.
- the V 2 O 5 content of 1.6% at 800 °C stone coal fired blank obtained after NH 4 + concentration in the clinker 2mol / L, HCO 3 - concentration of lmol / L ammonium bicarbonate solution vanadium leaching, the leaching temperature At 120 ° C, the solid matter mass ratio of the leachate was 6 m 3 /t, and the leaching time was 3 h, and the leaching residue and the ammonia-impregnated vanadium-containing liquid were obtained.
- the leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue. The recovery of vanadium in the vanadium slag was 88.2%.
- the clinker obtained by calcination of stone coal with a V 2 O 5 content of 1.2% at 950 ° C is leached in a solution of NH 4 + 6 mol/L, HCO 3 - 6 mol/L ammonium carbonate solution, and the leaching temperature is obtained.
- the leaching solution has a solid ratio of 5 m 3 /t and a leaching time of 2 h, and a leaching slag and an ammonia leaching vanadium containing solution are obtained.
- the leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue. The recovery of vanadium in the stone coal was 86.2%.
- the V 2 O 5 content of 7.5% at 700 °C vanadium-containing catalyst after calcination blank obtained clinker NH 4 + concentration 0.5mol / L, HCO 3 - concentration of 0.5mol / L ammonium bicarbonate solution leaching vanadium
- the leaching temperature was 60 ° C
- the leaching solution solid ratio was 7 m 3 /t
- the leaching time was 4 h
- the leaching slag and the ammonia leaching vanadium containing solution were obtained.
- the leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue.
- the recovery of vanadium in the vanadium slag was 89.1%.
- the clinker obtained by blank calcination of vanadium-containing catalyst with V 2 O 5 content of 8.3% at 800 ° C is at a concentration of 5 mol/L of NH 4 + , a concentration of 2 mol/L of HCO 3 and a concentration of 2 mol/L of CO 3 2- .
- the vanadium solution is leached to extract vanadium, the leaching temperature is 75 ° C, the leaching liquid solid ratio is 10 m 3 /t, and the leaching time is 2 h, and the leaching slag and the ammonia leaching vanadium containing solution are obtained.
- the leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue. The recovery of vanadium in the vanadium slag was 90.1%.
- the V 2 O 5 content of 1.5% vanadium titanium magnetite at 850 °C obtained after calcination clinker blank NH 4 + concentration 2mol / L, HCO 3 - concentration of 2mol / L ammonium bicarbonate solution leaching vanadium
- the leaching temperature is 35 ° C
- the leaching liquid solid ratio is 6 m 3 /t
- the leaching time is 1 h
- the leaching slag and the ammonia leaching vanadium containing solution are obtained.
- the leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue.
- the recovery of vanadium in the vanadium slag was 92.1%.
- the present invention illustrates the detailed process equipment and process flow of the present invention by the above embodiments, but the present invention is not limited to the above detailed process equipment and process flow, that is, does not mean that the present invention must rely on the above detailed process equipment and The process can only be implemented. It should be apparent to those skilled in the art that any modifications of the present invention, equivalent substitution of the various materials of the products of the present invention, addition of auxiliary components, selection of specific means, and the like, are all within the scope of the present invention.
Abstract
A method for extracting vanadium by leaching vanadium-containing raw material fired clinkers with an ammonium bicarbonate solution, the method comprising: leaching vanadium-containing raw material fired clinkers with an ammonium bicarbonate solution to enable vanadium to enter the solution in the form of ammonium vanadate to obtain leached residue and a vanadium-containing solution; and a vanadium pentoxide product is obtained from the vanadium-containing solution via cooling, crystallization and separation of the ammonium vanadate and calcination of the ammonium vanadate. The method employs an ammonium bicarbonate solution leaching technology without sodium salts, has a simple process, is easy to operate, and has a vanadium extraction rate of 85-99% without producing saline waste water.
Description
本发明属于钒化工冶金技术领域,尤其涉及一种含钒原料焙烧后熟料碳铵溶液浸出提钒的方法。The invention belongs to the technical field of vanadium chemical metallurgy, and in particular relates to a method for extracting vanadium from a clinker carbon ammonium solution after calcination of a vanadium-containing raw material.
钒钛磁铁矿、钒渣、石煤等含钒原料传统的提钒方法为焙烧-水浸/酸浸法提钒,其中含钒原料钠化焙烧-水浸是提钒的主流方法。钠化焙烧工艺的基本原理是以Na2CO3、NaCl、Na2SO4等钠盐为添加剂,通过高温焙烧(750-850℃)将含钒原料中低价态的钒转化为水溶性五价钒的钠盐,再对钠化焙烧产物直接水浸,得到含钒的浸取液,后加入铵盐制得多钒酸铵沉淀,经还原焙烧后获得钒的氧化物产品。钠化焙烧工艺钒回收率低,单次焙烧钒回收率为70%左右,经多次焙烧后钒的回收率也仅为80%,且需多次焙烧,能耗偏高;在焙烧过程中会产生有害的HCl、Cl2等侵蚀性气体,后续氨沉过程会得到高盐度含氨氮废水,含氨氮废水不仅污染环境,而且治理代价较高。Vanadium-titanium magnetite, vanadium slag, stone coal and other vanadium-containing raw materials The traditional vanadium extraction method is vanadium extraction by roasting-water immersion/acid leaching method. The vanadium-containing raw material sodium roasting-water leaching is the mainstream method for extracting vanadium. The basic principle of the sodium roasting process is to convert the low-valent vanadium in the vanadium-containing raw material into water-soluble five by high-temperature calcination (750-850 ° C) with sodium salt such as Na 2 CO 3 , NaCl, Na 2 SO 4 as an additive. The sodium salt of vanadium is directly immersed in the sodium calcined product to obtain a vanadium-containing leach solution, and then an ammonium salt is added to prepare ammonium vanadate precipitation, and the vanadium oxide product is obtained by reduction roasting. The recovery rate of vanadium in the sodium roasting process is low. The recovery rate of vanadium in a single roasting is about 70%. After repeated roasting, the recovery rate of vanadium is only 80%, and it needs to be calcined several times, and the energy consumption is high. During the roasting process. It will produce harmful HCl, Cl 2 and other aggressive gases. The subsequent ammonia precipitation process will result in high salinity ammonia-containing wastewater. The ammonia-containing wastewater not only pollutes the environment, but also has a high cost of treatment.
为解决钠化焙烧过程存在的废气污染,及因钠盐熔点低而产生的炉料结块结圈等问题,钙化焙烧-酸浸提钒作为一种较清洁的提钒工艺,受到了广泛关注。该方法是使用石灰、石灰石、白云石等石灰质原料替代钠盐添加剂进行钒渣高温焙烧,使钒转化为不溶于水的钒酸钙,再用稀酸浸出,浸出液中的钒用水解沉钒法或离子交换法回收。该方法曾在苏联图拉黑色冶金厂进行了工业应用,近年来攀钢集团(专利CN 101161831A)又在产品除杂与净化方面取得进展,并完成了中试。钙化焙烧可避免传统钠化焙烧的氯/硫污染,但由于同样受限于焙烧过程的传质障碍,钙化焙烧的钒转化率仍不足80%。
In order to solve the problem of exhaust gas pollution in the sodium roasting process and the agglomeration of the charge due to the low melting point of the sodium salt, calcification roasting-acid leaching vanadium has been widely concerned as a relatively clean vanadium extraction process. The method uses lime, limestone, dolomite and other calcareous raw materials instead of sodium salt additives to carry out high temperature roasting of vanadium slag, convert vanadium into water-insoluble calcium vanadate, and then dilute with dilute acid, and vanadium in the leachate is hydrolyzed by vanadium. Or ion exchange recovery. This method has been applied industrially in the Tula ferrous metallurgical plant in the Soviet Union. In recent years, Pangang Group (patent CN 101161831A) has made progress in product impurity removal and purification, and completed the pilot test. Calcification roasting avoids chlorine/sulfur contamination of conventional sodium roasting, but vanadium conversion for calcification roasting is still less than 80% due to mass transfer barriers that are also limited by the calcination process.
CN 101899582A提出一种由钒渣提取五氧化二钒的方法,钒渣于800~1000℃焙烧,熟料经过20~50%碱浓度的碱液碱浸,经过除硅、冷却结晶得到正钒酸钠晶体,后经过转溶、酸性铵盐沉钒得到钒酸铵。该方法存在杂质难以脱除、过滤困难的问题,同时避免不了酸性氨氮废水的问题,且其钒回收率低于90%。CN 101899582A proposes a method for extracting vanadium pentoxide from vanadium slag. The vanadium slag is calcined at 800-1000 ° C, and the clinker is subjected to alkali leaching of 20 to 50% alkali concentration, and the vanadic acid is obtained by silicon removal and cooling crystallization. The sodium crystals are then converted to vanadium by vanadium, which is converted to acid and ammonium salts. The method has the problem that impurities are difficult to remove and difficult to filter, and the problem of acid ammonia nitrogen wastewater can not be avoided, and the vanadium recovery rate is less than 90%.
CN 1082617A提出了对吹炼得到的高温钒渣在900~1300℃直接吹氧进行处理,促使渣中的低价钒氧化成为五价钒,渣冷却破碎后,在一定的温度、碱浓度、氧分压下浸出渣中的钒,该方法不必在钒渣降温后再次高温焙烧,能耗大为降低,并且避免了钠化焙烧造成的环境污染,但存在金属铁无法回收,且钒回收率低的问题。CN 1082617A proposes to treat the high-temperature vanadium slag obtained by blowing directly at 900-1300 ° C to promote the oxidation of low-valent vanadium in the slag into pentavalent vanadium. After the slag is cooled and crushed, at a certain temperature, alkali concentration, oxygen The vanadium in the slag is leached under partial pressure. The method does not need to be calcined at a high temperature after the vanadium slag is cooled down, the energy consumption is greatly reduced, and the environmental pollution caused by the sodium roasting is avoided, but the metal iron cannot be recovered, and the vanadium recovery rate is low. The problem.
目前未见有使用碳铵溶液直接浸出含钒原料焙烧熟料的相关报道。At present, there is no report on the direct leaching of vanadium-containing raw materials for roasting clinker using ammonium bicarbonate solution.
发明内容Summary of the invention
针对现有技术的不足,本发明的目的之一在于提供一种含钒原料焙烧后熟料碳铵溶液浸出提钒的方法。相对于现有的焙烧熟料浸出方法,该方法钒浸出率高,浸出温度低,浸出条件温和,且含钒原料中其它杂质不进入浸出液,后续钒酸铵分离简单。焙烧过程因不使用钠盐,操作易于控制,不产生有害窑气,且无含盐废水产生。In view of the deficiencies of the prior art, one of the objects of the present invention is to provide a method for leaching vanadium from a clinker-containing carbon ammonium solution after calcination of a vanadium-containing raw material. Compared with the existing roasting clinker leaching method, the vanadium leaching rate is high, the leaching temperature is low, the leaching conditions are mild, and other impurities in the vanadium-containing raw material do not enter the leachate, and subsequent ammonium vanadate separation is simple. Since the roasting process does not use sodium salt, the operation is easy to control, no harmful kiln gas is generated, and no salty wastewater is produced.
为达上述目的,本发明通过以下技术方案实现:To achieve the above object, the present invention is achieved by the following technical solutions:
一种含钒原料焙烧后熟料碳铵溶液浸出提钒的方法,将含钒原料焙烧后熟料在碳铵溶液中浸出提钒,使钒以钒酸铵形式进入溶液,得到浸出渣和含钒液。含钒原料焙烧后其中的三价钒在焙烧过程转化为可浸出的五价钒,五价钒可与碳铵溶液反应生成钒酸铵,从而将钒从原料中提取出来。A method for leaching vanadium by clinker carbon ammonium solution after calcination of a vanadium-containing raw material, pulverizing the crust-containing raw material in a solution of vanadium in a solution of vanadium, and causing vanadium to enter the solution in the form of ammonium vanadate to obtain leaching slag and Vanadium solution. After the vanadium-containing raw material is calcined, the trivalent vanadium is converted into a leachable pentavalent vanadium during the roasting process, and the pentavalent vanadium can be reacted with the ammonium bicarbonate solution to form ammonium vanadate, thereby extracting the vanadium from the raw material.
作为优选技术方案,本发明所述的方法,所述含钒原料为钒钛磁铁矿、石
煤、钒渣或含钒催化剂中的1种或2种以上的混合物;其中含钒催化剂可以为使用前的催化剂也可以为使用后无法再次使用的催化剂;As a preferred technical solution, in the method of the present invention, the vanadium-containing raw material is vanadium-titanium magnetite or stone.
a mixture of one or more of coal, vanadium slag or a vanadium-containing catalyst; wherein the vanadium-containing catalyst may be a catalyst before use or a catalyst that cannot be reused after use;
作为优选技术方案,本发明所述的方法,所述焙烧为经空白焙烧(即无添加剂焙烧)、钙化焙烧中的1种或2种以上的组合。According to a preferred embodiment of the present invention, the calcination is one or a combination of two or more of blank calcination (ie, no additive calcination) and calcification calcination.
作为优选技术方案,本发明所述的方法,所述碳铵溶液指的是NH4
+/HCO3
-/OH-、NH4
+/CO3
2-/OH-或NH4
+/HCO3
-/CO3
2-/OH-溶液,意即可以是NH4HCO3溶液、(NH4)2CO3溶液、NH4HCO3与(NH4)2CO3的混合溶液、氨水与NH4HCO3的混合溶液、氨水与(NH4)2CO3的混合溶液、氨水与NH4HCO3及(NH4)2CO3的混合液,也可通过在氨水中通入CO2气体碳化后获得。As a preferred technical solution, in the method of the present invention, the ammonium bicarbonate solution refers to NH 4 + /HCO 3 - /OH - , NH 4 + /CO 3 2- /OH - or NH 4 + /HCO 3 - /CO 3 2- /OH - solution, which means that it can be NH 4 HCO 3 solution, (NH 4 ) 2 CO 3 solution, mixed solution of NH 4 HCO 3 and (NH 4 ) 2 CO 3 , ammonia water and NH 4 HCO A mixed solution of 3, a mixed solution of ammonia and (NH 4 ) 2 CO 3 , a mixture of ammonia and NH 4 HCO 3 and (NH 4 ) 2 CO 3 can also be obtained by carbonizing a CO 2 gas in an ammonia water. .
作为优选技术方案,本发明所述的方法,所述碳铵溶液的浓度以溶液中各离子的摩尔浓度计:NH4
+浓度为0.5~6mol/L,HCO3
-浓度为0.5~6mol/L,CO3
2-的浓度为0.5~5mol/L,可通过配加不同溶液组合得到。碳铵溶液中NH4
+/HCO3
-/CO3
2-浓度低于0.5mol/L,则钒的提取率很低,NH4
+/HCO3
-浓度高于6mol/L,CO3
2-的浓度高于5mol/L,浸出时钒的浸出率基本不变,若再提高碳铵溶液浓度会大大增加提钒的成本,因此,本发明选择碳铵溶液的浓度为:NH4
+浓度为0.5~6mol/L,HCO3
-浓度为0.5~6mol/L,CO3
2-的浓度为0.5~5mol/L。As a preferred technical solution, in the method of the present invention, the concentration of the ammonium carbonate solution is determined by the molar concentration of each ion in the solution: the concentration of NH 4 + is 0.5-6 mol/L, and the concentration of HCO 3 - is 0.5-6 mol/L. The concentration of CO 3 2- is 0.5-5 mol/L, which can be obtained by adding different solution combinations. When the concentration of NH 4 + /HCO 3 - /CO 3 2- in the ammonium carbonate solution is less than 0.5mol/L, the extraction rate of vanadium is very low, and the concentration of NH 4 + /HCO 3 - is higher than 6mol/L, CO 3 2- The concentration of vanadium is more than 5mol/L, and the leaching rate of vanadium is basically unchanged during leaching. If the concentration of ammonium carbonate solution is increased, the cost of vanadium extraction will be greatly increased. Therefore, the concentration of the ammonium carbonate solution in the present invention is: NH 4 + concentration is The concentration of 0.5 to 6 mol/L, HCO 3 - is 0.5 to 6 mol/L, and the concentration of CO 3 2- is 0.5 to 5 mol/L.
作为优选技术方案,本发明所述的方法,所述浸出的温度为常温~150℃,例如为25℃、35℃、50℃、80℃、95℃、120℃、150℃等,温度太低含钒原料焙烧后熟料中的钒浸出效率低,150℃时各种不同焙烧熟料钒的浸出率已经很高,再升高浸出温度没有意义,因此,本发明选择浸出温度为常温~150℃,优选为50~95℃,再优选为60~75℃。As a preferred technical solution, in the method of the present invention, the temperature of the leaching is from normal temperature to 150 ° C, for example, 25 ° C, 35 ° C, 50 ° C, 80 ° C, 95 ° C, 120 ° C, 150 ° C, etc., the temperature is too low. The vanadium leaching efficiency of the clinker-containing raw material after calcination is low, and the leaching rate of vanadium of various roasting clinker at 150 ° C is already high, and it is meaningless to raise the leaching temperature. Therefore, the leaching temperature of the present invention is selected from normal temperature to 150 °C. °C is preferably 50 to 95 ° C, and more preferably 60 to 75 ° C.
作为优选技术方案,本发明所述的方法,所述浸出的时间为0.5h以上,例如为1h、2h、3h、4h等,浸出时间太短如低于0.5h,钒的提取效率低,优选为
1~3h。As a preferred technical solution, in the method of the present invention, the leaching time is 0.5 h or more, for example, 1 h, 2 h, 3 h, 4 h, etc., and the leaching time is too short, such as less than 0.5 h, and the extraction efficiency of vanadium is low, preferably. For
1 to 3 hours.
作为优选技术方案,本发明所述的方法,所述浸出的过程中碳铵溶液与含钒原料焙烧后熟料的液固体积质量比为2∶1~10∶1m3/t,例如为4∶1m3/t、7∶1m3/t、9∶1m3/t等,液固比在2m3/t以下反应液固传质差,难以实现,液固比在10m3/t浸出液中钒溶解量足够,提高液固比已再没有意义,优选为3∶1~6∶1m3/t。According to a preferred embodiment of the present invention, in the leaching process, the mass ratio of the liquid solid product of the ammonium carbonate solution to the clinker-containing raw material after calcination is 2:1 to 10:1 m 3 /t, for example, 4 : 1m 3 /t, 7:1m 3 /t, 9:1m 3 /t, etc., the liquid-solid ratio of the reaction liquid below 2m 3 /t is poor, and the liquid-solid ratio is 10m 3 /t. The amount of vanadium dissolved is sufficient, and it is no longer meaningful to increase the liquid-solid ratio, and is preferably from 3:1 to 6:1 m 3 /t.
作为优选技术方案,本发明所述的方法,包括如下步骤:将含钒原料焙烧后熟料在碳铵溶液中浸出提钒,碳铵溶液浓度为:NH4
+浓度0.5~6mol/L,HCO3
-浓度0.5~6mol/L,CO3
2-的浓度0.5~5mol/L,浸出温度为常温~150℃,浸出时间为0.5h以上,浸出过程中碳铵溶液与含钒原料焙烧后熟料的液固比为2∶1~10∶1m3/t,浸出后使钒以钒酸铵形式进入溶液,得到浸出渣和含钒液。As a preferred technical solution, the method of the present invention comprises the steps of: leaching vanadium in a solution of vanadium containing raw materials in a solution of ammonium carbonate, the concentration of the ammonium carbonate solution is: NH 4 + concentration of 0.5-6 mol/L, HCO 3 - concentration 0.5 ~ 6mol / L, CO 3 2- concentration 0.5 ~ 5mol / L, leaching temperature is normal temperature ~ 150 ° C, leaching time is more than 0.5h, carbon ammonium solution and vanadium containing raw materials after roasting clinker The liquid-solid ratio is from 2:1 to 10:1 m 3 /t. After leaching, vanadium is introduced into the solution as ammonium vanadate to obtain a leaching residue and a vanadium-containing solution.
本发明相比传统的焙烧熟料水浸/酸浸工艺,钒浸取率高,且杂质不进入含钒液,可实现钒的简单、清洁分离。Compared with the conventional roasting clinker water immersion/acid leaching process, the vanadium leaching rate is high, and impurities do not enter the vanadium-containing liquid, and the simple and clean separation of vanadium can be realized.
为便于理解本发明,本发明列举实施例如下。本领域技术人员应该明了,所述实施例仅仅是帮助理解本发明,不应视为对本发明的具体限制。To facilitate an understanding of the invention, the invention is set forth below. It should be understood by those skilled in the art that the present invention is not to be construed as limited.
实施例1Example 1
将V2O5含量为11.6%的钒渣在900℃经空白焙烧后得到的熟料在NH4
+浓度6mol/L,HCO3
-浓度6mol/L的碳铵溶液中浸出提钒,浸出温度为95℃,浸出液固体积质量比为10m3/t,浸出时间为3h,得到浸出渣和含钒液。浸出渣经洗涤、烘干、称重并分析残渣的钒含量,钒渣中钒的回收率为94.2%。The clinker obtained by blank calcination of vanadium slag with a V 2 O 5 content of 11.6% at 900 ° C was leached in a solution of NH 4 + 6 mol/L, HCO 3 - 6 mol/L ammonium nitrate solution, and the leaching temperature was obtained. At 95 ° C, the solid solution mass ratio of the leachate was 10 m 3 /t, and the leaching time was 3 h, and the leaching residue and the vanadium-containing liquid were obtained. The leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue. The recovery of vanadium in the vanadium slag was 94.2%.
实施例2Example 2
将V2O5含量为1.5%的钒钛磁铁矿在850℃经钙化焙烧后得到的熟料在NH4
+浓度4mol/L,HCO3
-浓度1mol/L,CO3
2-的浓度1mol/L的碳铵溶液中浸出
提钒,浸出温度为75℃,浸出液固体积质量比为8m3/t,浸出时间为2h,得到浸出渣和含钒液。浸出渣经洗涤、烘干、称重并分析残渣的钒含量,钒渣中钒的回收率为90.2%。The V 2 O 5 content of 1.5% vanadium titanium magnetite at 850 ℃ obtained after roasting clinker NH 4 + concentration 4mol / L, HCO 3 -, CO 3 2- concentration concentration 1mol / L 1mol The vanadium extract was extracted from the ammonium carbonate solution of /L, the leaching temperature was 75 ° C, the solid solution mass ratio of the leachate was 8 m 3 /t, and the leaching time was 2 h, and the leaching residue and the vanadium-containing solution were obtained. The leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue. The recovery of vanadium in the vanadium slag was 90.2%.
实施例3Example 3
将V2O5含量为14.7%的钒渣在800℃经钙化焙烧后得到的熟料在NH4
+浓度5mol/L,HCO3
-浓度2mol/L,CO3
2-的浓度lmol/L的碳铵溶液中浸出提钒,浸出温度为150℃,浸出液固体积质量比为5m3/t,浸出时间为2.5h,得到浸出渣和氨浸含钒液。浸出渣经洗涤、烘干、称重并分析残渣的钒含量,钒渣中钒的回收率为93.4%。The clinker obtained by calcination of vanadium slag with a V 2 O 5 content of 14.7% at 800 ° C is at a concentration of 5 mol/L of NH 4 + , a concentration of 2 mol/L of HCO 3 and a concentration of 1 mol/L of CO 3 2- . The vanadium extraction solution was leached with vanadium, the leaching temperature was 150 ° C, the solid matter mass ratio of the leaching solution was 5 m 3 /t, and the leaching time was 2.5 h, and the leaching residue and the ammonia leaching vanadium containing solution were obtained. The leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue. The recovery of vanadium in the vanadium slag was 93.4%.
实施例4Example 4
将V2O5含量为1.6%的石煤在800℃经空白焙烧后得到的熟料在NH4
+浓度2mol/L,HCO3
-浓度lmol/L的碳铵溶液中浸出提钒,浸出温度为120℃,浸出液固体积质量比为6m3/t,浸出时间为3h,得到浸出渣和氨浸含钒液。浸出渣经洗涤、烘干、称重并分析残渣的钒含量,钒渣中钒的回收率为88.2%。The V 2 O 5 content of 1.6% at 800 ℃ stone coal fired blank obtained after NH 4 + concentration in the clinker 2mol / L, HCO 3 - concentration of lmol / L ammonium bicarbonate solution vanadium leaching, the leaching temperature At 120 ° C, the solid matter mass ratio of the leachate was 6 m 3 /t, and the leaching time was 3 h, and the leaching residue and the ammonia-impregnated vanadium-containing liquid were obtained. The leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue. The recovery of vanadium in the vanadium slag was 88.2%.
实施例5Example 5
将V2O5含量为1.2%的石煤在950℃经钙化焙烧后得到的熟料在NH4
+浓度6mol/L,HCO3
-浓度6mol/L的碳铵溶液中浸出提钒,浸出温度为85℃,浸出液固比为5m3/t,浸出时间为2h,得到浸出渣和氨浸含钒液。浸出渣经洗涤、烘干、称重并分析残渣的钒含量,石煤中钒的回收率为86.2%。The clinker obtained by calcination of stone coal with a V 2 O 5 content of 1.2% at 950 ° C is leached in a solution of NH 4 + 6 mol/L, HCO 3 - 6 mol/L ammonium carbonate solution, and the leaching temperature is obtained. At 85 ° C, the leaching solution has a solid ratio of 5 m 3 /t and a leaching time of 2 h, and a leaching slag and an ammonia leaching vanadium containing solution are obtained. The leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue. The recovery of vanadium in the stone coal was 86.2%.
实施例6Example 6
将V2O5含量为7.5%的含钒催化剂在700℃经空白焙烧后得到的熟料在NH4
+浓度0.5mol/L,HCO3
-浓度0.5mol/L的碳铵溶液中浸出提钒,浸出温度为60℃,浸出液固比为7m3/t,浸出时间为4h,得到浸出渣和氨浸含钒液。浸出
渣经洗涤、烘干、称重并分析残渣的钒含量,钒渣中钒的回收率为89.1%。The V 2 O 5 content of 7.5% at 700 ℃ vanadium-containing catalyst after calcination blank obtained clinker NH 4 + concentration 0.5mol / L, HCO 3 - concentration of 0.5mol / L ammonium bicarbonate solution leaching vanadium The leaching temperature was 60 ° C, the leaching solution solid ratio was 7 m 3 /t, and the leaching time was 4 h, and the leaching slag and the ammonia leaching vanadium containing solution were obtained. The leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue. The recovery of vanadium in the vanadium slag was 89.1%.
实施例7Example 7
将V2O5含量为8.3%的含钒催化剂在800℃经空白焙烧后得到的熟料在NH4
+浓度5mol/L,HCO3
-浓度2mol/L,CO3
2-的浓度2mol/L的碳铵溶液中浸出提钒,浸出温度为75℃,浸出液固比为10m3/t,浸出时间为2h,得到浸出渣和氨浸含钒液。浸出渣经洗涤、烘干、称重并分析残渣的钒含量,钒渣中钒的回收率为90.1%。The clinker obtained by blank calcination of vanadium-containing catalyst with V 2 O 5 content of 8.3% at 800 ° C is at a concentration of 5 mol/L of NH 4 + , a concentration of 2 mol/L of HCO 3 and a concentration of 2 mol/L of CO 3 2- . The vanadium solution is leached to extract vanadium, the leaching temperature is 75 ° C, the leaching liquid solid ratio is 10 m 3 /t, and the leaching time is 2 h, and the leaching slag and the ammonia leaching vanadium containing solution are obtained. The leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue. The recovery of vanadium in the vanadium slag was 90.1%.
实施例8Example 8
将V2O5含量为1.5%的钒钛磁铁矿在850℃经空白焙烧后得到的熟料在NH4
+浓度2mol/L,HCO3
-浓度2mol/L的碳铵溶液中浸出提钒,浸出温度为35℃,浸出液固比为6m3/t,浸出时间为1h,得到浸出渣和氨浸含钒液。浸出渣经洗涤、烘干、称重并分析残渣的钒含量,钒渣中钒的回收率为92.1%。The V 2 O 5 content of 1.5% vanadium titanium magnetite at 850 ℃ obtained after calcination clinker blank NH 4 + concentration 2mol / L, HCO 3 - concentration of 2mol / L ammonium bicarbonate solution leaching vanadium The leaching temperature is 35 ° C, the leaching liquid solid ratio is 6 m 3 /t, and the leaching time is 1 h, and the leaching slag and the ammonia leaching vanadium containing solution are obtained. The leaching residue was washed, dried, weighed and analyzed for the vanadium content of the residue. The recovery of vanadium in the vanadium slag was 92.1%.
申请人声明,本发明通过上述实施例来说明本发明的详细工艺设备和工艺流程,但本发明并不局限于上述详细工艺设备和工艺流程,即不意味着本发明必须依赖上述详细工艺设备和工艺流程才能实施。所属技术领域的技术人员应该明了,对本发明的任何改进,对本发明产品各原料的等效替换及辅助成分的添加、具体方式的选择等,均落在本发明的保护范围和公开范围之内。
The Applicant declares that the present invention illustrates the detailed process equipment and process flow of the present invention by the above embodiments, but the present invention is not limited to the above detailed process equipment and process flow, that is, does not mean that the present invention must rely on the above detailed process equipment and The process can only be implemented. It should be apparent to those skilled in the art that any modifications of the present invention, equivalent substitution of the various materials of the products of the present invention, addition of auxiliary components, selection of specific means, and the like, are all within the scope of the present invention.
Claims (9)
- 一种含钒原料焙烧后熟料碳铵溶液浸出提钒的方法,其特征在于,将含钒原料焙烧后熟料在碳铵溶液中浸出提钒,使钒以钒酸铵形式进入溶液,得到浸出渣和含钒液。A method for extracting vanadium by clinker carbon ammonium solution after calcination of a vanadium-containing raw material, characterized in that: after calcining the vanadium-containing raw material, the clinker is leached in a solution of ammonium carbonate to extract vanadium, and vanadium is introduced into the solution in the form of ammonium vanadate. Leaching slag and vanadium containing solution.
- 根据权利要求1所述的方法,其特征在于,所述含钒原料为钒钛磁铁矿、石煤、钒渣或含钒催化剂中的1种或2种以上的混合物;其中含钒催化剂为使用前的催化剂或为使用后无法再利用的催化剂。The method according to claim 1, wherein the vanadium-containing material is one or a mixture of two or more of vanadium-titanium magnetite, stone coal, vanadium slag or a vanadium-containing catalyst; wherein the vanadium-containing catalyst is The catalyst before use or the catalyst that cannot be reused after use.
- 根据权利要求1或2所述的方法,其特征在于,所述焙烧为空白焙烧、钙化焙烧中的1种或2种以上的组合。The method according to claim 1 or 2, wherein the calcination is one or a combination of two or more of blank calcination and calcification calcination.
- 根据权利要求1-3任一项所述的方法,其特征在于,所述碳铵溶液是NH4HCO3溶液、(NH4)2CO3溶液、NH4HCO3与(NH4)2CO3的混合溶液、氨水与NH4HCO3的混合溶液、氨水与(NH4)2CO3的混合溶液、氨水与NH4HCO3及(NH4)2CO3的混合液。The method according to any one of claims 1 to 3, wherein the ammonium bicarbonate solution is a NH 4 HCO 3 solution, a (NH 4 ) 2 CO 3 solution, NH 4 HCO 3 and (NH 4 ) 2 CO A mixed solution of 3, a mixed solution of ammonia water and NH 4 HCO 3 , a mixed solution of ammonia water and (NH 4 ) 2 CO 3 , a mixed solution of ammonia water and NH 4 HCO 3 and (NH 4 ) 2 CO 3 .
- 根据权利要求1-4任一项所述的方法,其特征在于,所述碳铵溶液的浓度以溶液中各离子的摩尔浓度计:NH4 +浓度为0.5~6mol/L,HCO3 -浓度为0.5~6mol/L,CO3 2-的浓度为0.5~5mol/L。The method according to any one of claims 1 to 4, wherein the concentration of the ammonium carbonate solution is based on a molar concentration of each ion in the solution: a concentration of NH 4 + of 0.5 to 6 mol/L, and a concentration of HCO 3 - It is 0.5 to 6 mol/L, and the concentration of CO 3 2- is 0.5 to 5 mol/L.
- 根据权利要求1-5任一项所述的方法,其特征在于,所述的浸出温度为常温~150℃,优选为50~95℃,再优选为60~75℃。The method according to any one of claims 1 to 5, wherein the leaching temperature is from ordinary temperature to 150 ° C, preferably from 50 to 95 ° C, and more preferably from 60 to 75 ° C.
- 根据权利要求1-6任一项所述的方法,其特征在于,所述的浸出时间为0.5h以上,优选为1~3h。The method according to any one of claims 1 to 6, wherein the leaching time is 0.5 h or more, preferably 1 to 3 h.
- 根据权利要求1-7任一项所述的方法,其特征在于,所述浸出过程中碳铵溶液与含钒原料焙烧后熟料的液固体积质量比为2∶1~10∶1m3/t,优选为3∶1~6∶1m3/t。The method according to any one of claims 1 to 7, wherein the mass ratio of the liquid solid product of the ammonium carbonate solution to the clinker-containing raw material after the leaching process is 2:1 to 10:1 m 3 / t is preferably from 3:1 to 6:1 m 3 /t.
- 根据权利要求1所述的方法,其特征在于,包括如下步骤:将含钒原料 焙烧后熟料在碳铵溶液中浸出提钒,碳铵溶液浓度为:NH4 +浓度0.5~6mol/L,HCO3 -浓度0.5~6mol/L,CO3 2-的浓度0.5~5mol/L,浸出温度为常温~150℃,浸出时间为0.5h以上,浸出过程中碳铵溶液与含钒原料焙烧后熟料的液固比为2∶1~10∶1m3/t,浸出后使钒以钒酸铵形式进入溶液,得到浸出渣和含钒液。 The method according to claim 1, comprising the steps of: leaching vanadium in the ammonium carbonate solution after calcining the vanadium-containing raw material, wherein the concentration of the ammonium carbonate solution is: NH 4 + concentration of 0.5 to 6 mol/L, HCO 3 - concentration of 0.5 ~ 6mol / L, CO 3 2- concentration of 0.5 ~ 5mol / L, leaching temperature is room temperature ~ 150 ℃, leaching time of 0.5h or more, during leaching solution containing ammonium bicarbonate baking cooked vanadium material The liquid-solid ratio of the material is 2:1 to 10:1 m3/t. After leaching, vanadium is introduced into the solution as ammonium vanadate to obtain leaching slag and vanadium-containing liquid.
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