CN115821078A - Method for cooperatively treating fluorite concentrate and iron tailings - Google Patents
Method for cooperatively treating fluorite concentrate and iron tailings Download PDFInfo
- Publication number
- CN115821078A CN115821078A CN202211522585.4A CN202211522585A CN115821078A CN 115821078 A CN115821078 A CN 115821078A CN 202211522585 A CN202211522585 A CN 202211522585A CN 115821078 A CN115821078 A CN 115821078A
- Authority
- CN
- China
- Prior art keywords
- leaching
- sulfuric acid
- rare earth
- niobium
- slag
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Pending
Links
- XEEYBQQBJWHFJM-UHFFFAOYSA-N Iron Chemical compound [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 title claims abstract description 108
- WUKWITHWXAAZEY-UHFFFAOYSA-L calcium difluoride Chemical compound [F-].[F-].[Ca+2] WUKWITHWXAAZEY-UHFFFAOYSA-L 0.000 title claims abstract description 57
- 239000010436 fluorite Substances 0.000 title claims abstract description 57
- 229910052742 iron Inorganic materials 0.000 title claims abstract description 54
- 239000012141 concentrate Substances 0.000 title claims abstract description 49
- 238000000034 method Methods 0.000 title claims abstract description 44
- 238000002386 leaching Methods 0.000 claims abstract description 82
- 229910052761 rare earth metal Inorganic materials 0.000 claims abstract description 59
- 239000010955 niobium Substances 0.000 claims abstract description 57
- 150000002910 rare earth metals Chemical class 0.000 claims abstract description 57
- 229910052758 niobium Inorganic materials 0.000 claims abstract description 54
- GUCVJGMIXFAOAE-UHFFFAOYSA-N niobium atom Chemical compound [Nb] GUCVJGMIXFAOAE-UHFFFAOYSA-N 0.000 claims abstract description 54
- 239000002893 slag Substances 0.000 claims abstract description 46
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims abstract description 39
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims abstract description 38
- 229910001868 water Inorganic materials 0.000 claims abstract description 38
- 239000002245 particle Substances 0.000 claims abstract description 28
- 238000002156 mixing Methods 0.000 claims abstract description 16
- 238000001035 drying Methods 0.000 claims abstract description 15
- 238000001914 filtration Methods 0.000 claims abstract description 11
- HIFJUMGIHIZEPX-UHFFFAOYSA-N sulfuric acid;sulfur trioxide Chemical compound O=S(=O)=O.OS(O)(=O)=O HIFJUMGIHIZEPX-UHFFFAOYSA-N 0.000 claims description 16
- 238000006243 chemical reaction Methods 0.000 claims description 14
- KRHYYFGTRYWZRS-UHFFFAOYSA-N Fluorane Chemical compound F KRHYYFGTRYWZRS-UHFFFAOYSA-N 0.000 claims description 12
- OSGAYBCDTDRGGQ-UHFFFAOYSA-L calcium sulfate Chemical compound [Ca+2].[O-]S([O-])(=O)=O OSGAYBCDTDRGGQ-UHFFFAOYSA-L 0.000 claims description 12
- 239000000395 magnesium oxide Substances 0.000 claims description 11
- CPLXHLVBOLITMK-UHFFFAOYSA-N magnesium oxide Inorganic materials [Mg]=O CPLXHLVBOLITMK-UHFFFAOYSA-N 0.000 claims description 11
- AXZKOIWUVFPNLO-UHFFFAOYSA-N magnesium;oxygen(2-) Chemical compound [O-2].[Mg+2] AXZKOIWUVFPNLO-UHFFFAOYSA-N 0.000 claims description 11
- 229910000040 hydrogen fluoride Inorganic materials 0.000 claims description 10
- 239000000203 mixture Substances 0.000 claims description 8
- 239000007789 gas Substances 0.000 claims description 7
- 238000007885 magnetic separation Methods 0.000 claims description 7
- 239000006148 magnetic separator Substances 0.000 claims description 7
- 230000003472 neutralizing effect Effects 0.000 claims description 7
- 238000001354 calcination Methods 0.000 claims description 6
- 230000035484 reaction time Effects 0.000 claims description 3
- 230000002195 synergetic effect Effects 0.000 abstract description 4
- 230000000052 comparative effect Effects 0.000 description 9
- 229910052602 gypsum Inorganic materials 0.000 description 9
- 239000010440 gypsum Substances 0.000 description 9
- 230000009286 beneficial effect Effects 0.000 description 7
- 229910052500 inorganic mineral Inorganic materials 0.000 description 6
- 239000011707 mineral Substances 0.000 description 6
- ZSLUVFAKFWKJRC-IGMARMGPSA-N 232Th Chemical compound [232Th] ZSLUVFAKFWKJRC-IGMARMGPSA-N 0.000 description 5
- 229910052776 Thorium Inorganic materials 0.000 description 5
- 238000000354 decomposition reaction Methods 0.000 description 5
- 239000012535 impurity Substances 0.000 description 5
- 239000000843 powder Substances 0.000 description 5
- FAPWRFPIFSIZLT-UHFFFAOYSA-M Sodium chloride Chemical compound [Na+].[Cl-] FAPWRFPIFSIZLT-UHFFFAOYSA-M 0.000 description 4
- 239000002253 acid Substances 0.000 description 4
- 229910052751 metal Inorganic materials 0.000 description 4
- 239000002184 metal Substances 0.000 description 4
- 150000002739 metals Chemical class 0.000 description 4
- 238000006386 neutralization reaction Methods 0.000 description 4
- 238000012545 processing Methods 0.000 description 4
- 238000012360 testing method Methods 0.000 description 4
- 239000000463 material Substances 0.000 description 3
- 239000008188 pellet Substances 0.000 description 3
- 239000002994 raw material Substances 0.000 description 3
- 238000011084 recovery Methods 0.000 description 3
- 229910004261 CaF 2 Inorganic materials 0.000 description 2
- KCXVZYZYPLLWCC-UHFFFAOYSA-N EDTA Chemical compound OC(=O)CN(CC(O)=O)CCN(CC(O)=O)CC(O)=O KCXVZYZYPLLWCC-UHFFFAOYSA-N 0.000 description 2
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 description 2
- 238000004458 analytical method Methods 0.000 description 2
- 238000000498 ball milling Methods 0.000 description 2
- TZCXTZWJZNENPQ-UHFFFAOYSA-L barium sulfate Chemical compound [Ba+2].[O-]S([O-])(=O)=O TZCXTZWJZNENPQ-UHFFFAOYSA-L 0.000 description 2
- AXCZMVOFGPJBDE-UHFFFAOYSA-L calcium dihydroxide Chemical compound [OH-].[OH-].[Ca+2] AXCZMVOFGPJBDE-UHFFFAOYSA-L 0.000 description 2
- 239000000920 calcium hydroxide Substances 0.000 description 2
- 229910001861 calcium hydroxide Inorganic materials 0.000 description 2
- 229960001484 edetic acid Drugs 0.000 description 2
- 238000005188 flotation Methods 0.000 description 2
- 238000010438 heat treatment Methods 0.000 description 2
- KMUONIBRACKNSN-UHFFFAOYSA-N potassium dichromate Chemical compound [K+].[K+].[O-][Cr](=O)(=O)O[Cr]([O-])(=O)=O KMUONIBRACKNSN-UHFFFAOYSA-N 0.000 description 2
- 229910052706 scandium Inorganic materials 0.000 description 2
- SIXSYDAISGFNSX-UHFFFAOYSA-N scandium atom Chemical compound [Sc] SIXSYDAISGFNSX-UHFFFAOYSA-N 0.000 description 2
- 238000000926 separation method Methods 0.000 description 2
- 239000011780 sodium chloride Substances 0.000 description 2
- OAICVXFJPJFONN-UHFFFAOYSA-N Phosphorus Chemical compound [P] OAICVXFJPJFONN-UHFFFAOYSA-N 0.000 description 1
- 125000004122 cyclic group Chemical group 0.000 description 1
- 238000002474 experimental method Methods 0.000 description 1
- 230000002349 favourable effect Effects 0.000 description 1
- 239000010419 fine particle Substances 0.000 description 1
- 239000008187 granular material Substances 0.000 description 1
- 238000004519 manufacturing process Methods 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 239000011574 phosphorus Substances 0.000 description 1
- 229910052698 phosphorus Inorganic materials 0.000 description 1
- 238000001556 precipitation Methods 0.000 description 1
- 238000007670 refining Methods 0.000 description 1
- 238000011160 research Methods 0.000 description 1
- 239000007787 solid Substances 0.000 description 1
- 238000002798 spectrophotometry method Methods 0.000 description 1
- 238000006467 substitution reaction Methods 0.000 description 1
- 238000004448 titration Methods 0.000 description 1
- 238000005406 washing Methods 0.000 description 1
- 239000002699 waste material Substances 0.000 description 1
Classifications
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
Landscapes
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention discloses a method for cooperatively treating fluorite concentrate and iron tailings, which comprises the following steps: 1) Mixing fluorite concentrate and iron tailings in a mass ratio of 1 (0.1-0.4), granulating, and drying to obtain dried particles; 2) Pre-reacting the dried particles with sulfuric acid at 100-145 ℃, and then roasting at 210-270 ℃ to obtain roasting slag; 3) Leaching the roasting slag with water, and filtering to obtain leaching solution containing rare earth and niobium and water leaching slag. The method can perform synergistic treatment on the fluorite concentrate and the iron tailings, and has higher leaching rates of the rare earth and the niobium.
Description
Technical Field
The invention relates to a method for cooperatively treating fluorite concentrate and iron tailings.
Background
On the one hand, in order to obtain high-quality iron ore concentrate, a reverse flotation process is generally adopted to remove micro-fine particles containing iron intergrowths and inclusions, so that high-grade iron ore concentrate and lower-grade iron tailings are obtained. The obtained iron tailings contain higher valuable resources such as iron, rare earth, niobium, fluorite and the like. Because the embedding granularity is too fine, several minerals cannot be effectively separated by adopting a conventional mineral separation method, and the conventional mineral separation method mainly deposits the minerals, so that not only is the resource waste caused, but also the environment is polluted. On the other hand, fluorite concentrate obtained by flotation of fluorite by taking rare earth tailings as a raw material is poor in quality and low in utilization rate, and valuable resources such as rare earth and niobium are contained, so that the rare earth and niobium are not fully utilized.
Therefore, how to simultaneously recover valuable metals in the iron tailings and the fluorite concentrate and improve the leaching rate of the valuable metals has important significance.
CN109371239A discloses a method for processing rare earth-containing low-grade fluorite ore, which adopts a method of absolute excess sulfuric acid solution program temperature control subsection pulpifying decomposition processing, controls the low temperature section to decompose fluorite mineral preferentially, increases the temperature to decompose rare earth mineral, recovers hydrofluoric acid from tail gas, uses water to leach rare earth from acid leaching slag after reaction, neutralizes thorium, and recovers rare earth and gypsum. This patent document only mentions the REO decomposition rate and does not refer to the leaching rates of rare earth and niobium.
CN113735062A discloses a method for preparing hydrogen fluoride by rare earth recovery fluorite, which comprises processing fluorite into powder, refining the powder into high-purity acid-grade fluorite fine powder, and reacting the high-purity acid-grade fluorite fine powder with industrial sulfuric acid to prepare the high-purity hydrogen fluoride. The raw material adopts high-purity acid-grade fluorite fine powder, and although the fuming acid proportion is mentioned in the production process, specific numerical values and specific reaction temperature are not mentioned. More importantly, the recovery of rare earths and niobium is not considered.
CN106636614A discloses a method for leaching niobium, scandium and rare earth elements from tailings, comprising the following steps: s1, adding calcium hydroxide and sodium chloride into tailings of iron selection, rare earth selection and fluorite, uniformly mixing to obtain a mixture, and roasting the mixture to obtain roasted ore; s2, carrying out ball milling treatment on the roasted ore; s3, mixing the roasted ore subjected to ball milling with hydrochloric acid, heating for leaching, and filtering a leachate to obtain leaching residue I and a leachate I rich in rare earth and scandium; s4, drying the leaching residue I, leaching the leaching residue I by adopting a concentrated sulfuric acid heating leaching method, and filtering the leachate to obtain leaching residue II and a leaching solution II rich in niobium. In this patent document, calcium hydroxide and sodium chloride need to be added, and the calcination temperature is high, which is not suitable for industrial application.
Disclosure of Invention
In view of the above, the present invention is directed to a method for co-processing fluorite concentrate and iron tailings, which can co-process fluorite concentrate and iron tailings. Furthermore, the leaching rate of the rare earth and niobium is higher. The invention adopts the following technical scheme to achieve the purpose.
The invention provides a method for cooperatively treating fluorite concentrate and iron tailings, which comprises the following steps:
1) Mixing fluorite concentrate and iron tailings in a mass ratio of 1 (0.1-0.4), granulating, and drying to obtain dried particles; wherein CaF in the fluorite concentrate 2 More than 75wt%, REO 4.5-9 wt%, nb 2 O 5 The content is 0.07 to 0.5 weight percent, and the content of P is 0.5 to 1.4 weight percent; the iron tailings comprise 30 to 45 weight percent of TFe, 4.6 to 8 weight percent of REO and Nb 2 O 5 The content is 0.07-0.5 wt%, and the P content is 0.8-1.7 wt%;
2) Pre-reacting the dried particles with sulfuric acid at 100-145 ℃, and then roasting at 210-270 ℃ to obtain gas containing hydrogen fluoride and roasting slag;
3) Leaching the roasting slag with water, and filtering to obtain leaching solution containing rare earth and niobium and water leaching slag.
According to the method of the present invention, preferably, the dried particles are obtained by a process comprising the steps of:
mixing the fluorite concentrate with the iron tailings, preparing the mixture into balls, and drying the prepared balls for 20-38 hours at the temperature of 90-120 ℃ to obtain dried particles.
According to the process of the present invention, preferably, in step 2), the sulfuric acid is a mixture of 98wt% concentrated sulfuric acid and 105wt% fuming sulfuric acid.
According to the method of the invention, the mass ratio of the dried particles, 98wt% concentrated sulfuric acid and 105wt% fuming sulfuric acid is preferably 1 (0.85-1.0): (0.5-0.8).
According to the method of the present invention, preferably, in the step 2), the pre-reaction temperature is 100 to 135 ℃, and the pre-reaction time is 4 to 9min.
According to the method of the invention, preferably, in the step 2), the roasting temperature is 210-260 ℃ and the roasting time is 90-130 min.
According to the method of the invention, preferably, in the step 2), the roasting temperature is 220-260 ℃ and the roasting time is 100-130 min.
According to the method of the invention, preferably, in the step 3), the mass ratio of the roasting slag to the water is 1 (0.11-0.2).
The method according to the present invention preferably further comprises the steps of:
magnetically separating the water leaching slag obtained in the step 3) by using a magnetic separator to obtain calcium sulfate; wherein, the magnetic separation intensity is 3.0-4.2T.
The method according to the present invention preferably further comprises the steps of:
neutralizing the leaching solution containing the rare earth and the niobium obtained in the step 3) by adopting magnesium oxide to obtain neutralized leaching solution containing the rare earth and the niobium and neutralized slag.
The method can perform synergistic treatment on the fluorite concentrate and the iron tailings, can recover valuable metals of rare earth and niobium, and has high leaching rate of the rare earth and niobium. The method of the invention can also obtain gypsum with higher purity. In addition, anhydrous hydrogen fluoride can also be obtained. According to the preferred technical scheme, the leaching rate of the rare earth and the niobium is improved by synergistically treating a small amount of iron tailings and fluorite concentrate.
Detailed Description
The present invention will be further described with reference to the following specific examples, but the scope of the present invention is not limited thereto.
The method for the synergistic treatment of fluorite concentrate and iron tailings comprises the following steps: 1) Granulating and drying; 2) A pre-reaction step; 3) Roasting; 4) Leaching; 5) A magnetic separation step, and 6) a neutralization step. Wherein, the neutralization step and the magnetic separation step can be out of sequence. As described in detail below.
< granulating and drying step >
And mixing the fluorite concentrate and the iron tailings, granulating, and drying to obtain dried particles. According to the invention, the fluorite concentrate and the iron tailings are subjected to synergistic treatment, so that the environment is protected, valuable metals in the fluorite concentrate and the iron tailings can be recovered, and the leaching rate of rare earth and niobium is improved.
The content of TFe (total iron) in the iron tailings can be 30-45 wt%, preferably 34-42 wt%, more preferably 35-40 wt%, and still more preferably 36.8-37.3 wt%. The REO content may be 4.6 to 8wt%, preferably 5.1 to 7.5wt%, more preferably 5.9 to 7.5wt%, and still more preferably 7.25 to 7.4wt%. Nb 2 O 5 The content may be 0.07 to 0.5wt%, preferably 0.11 to 0.4wt%, more preferably 0.14 to 0.3wt%, and still more preferably 0.15 to 0.25wt%. The P content may be 0.8 to 1.7wt%, preferably 1.0 to 1.6wt%, more preferably 1.1 to 1.4wt%. The iron tailings also contain CaF 2 The content thereof may be 16 to 30wt%, preferably 18 to 28wt%, more preferably 19 to 28wt%.
CaF in fluorite concentrate 2 The content is more than 75wt%, preferably more than or equal to 78wt%, more preferably more than or equal to 80wt% and less than 92wt%, still more preferably more than or equal to 84wt% and less than 86wt%. The REO content may be 4.5 to 9wt%, preferably 4.8 to 8wt%, more preferably 5.5 to 7.0wt%, and still more preferably 5.85 to 6.3wt%. Nb 2 O 5 The content can be 0.07-0.5 wt%Preferably 0.09 to 0.4wt%, more preferably 0.11 to 0.25wt%, and still more preferably 0.19 to 0.21wt%. The P content (i.e., phosphorus content) is 0.5 to 1.4wt%, preferably 0.7 to 1.2wt%, more preferably 0.8 to 1.0wt%, and still more preferably 0.84 to 0.95wt%. Such fluorite concentrates and iron tailings can be processed in the particular process and process parameter ranges of the present invention with higher leaching rates of rare earths and niobium.
The mass ratio of the fluorite concentrate to the iron tailings can be 1 (0.1-0.4), preferably 1 (0.15-0.35), and more preferably 1 (0.2-0.3). Thus being beneficial to improving the leaching rate of the rare earth and the niobium.
The fluorite concentrate is mixed with the iron tailings and then granulated, preferably to form balls, and the diameter of the balls can be 2-5 mm, preferably 2-4 mm, and more preferably 3-4 mm. The ball in the range is beneficial to the reaction and the recovery of the dried particles. In the present invention, the pellets may be produced using a disk granulator.
The drying temperature may be 90 to 120 ℃, preferably 90 to 110 ℃, more preferably 95 to 100 ℃. The drying time may be 20 to 38 hours, preferably 24 to 36 hours, and more preferably 24 to 32 hours. The moisture content of the oven dried pellets is less than 0.1wt%. This is advantageous for improving the decomposition rate of rare earth and niobium.
According to one embodiment of the invention, the prepared pellets are dried at 90-100 ℃ for 24-30 h to obtain dried particles.
< Pre-reaction step >
And carrying out pre-reaction on the dried particles and sulfuric acid at the temperature of 100-145 ℃ to obtain a pre-reaction material. Thus being beneficial to improving the decomposition rate of the rare earth and the niobium and further being beneficial to improving the leaching rate of the rare earth and the niobium.
In the present invention, the sulfuric acid is a mixture of 98wt% concentrated sulfuric acid and 105wt% fuming sulfuric acid. The mass ratio of 98wt% concentrated sulfuric acid to 105wt% fuming sulfuric acid may be 0.85-1.0: 0.5 to 0.8, preferably 0.9 to 1.0, more preferably 0.9 to 1.0.
In the invention, the mass ratio of the dried particles to 98wt% concentrated sulfuric acid and 105wt% fuming sulfuric acid can be 1 (0.85-1.0): (0.5 to 0.8), preferably 1 (0.9 to 1.0): (0.6-0.8).
The pre-reaction temperature may be 100 to 145 ℃, preferably 100 to 135 ℃, more preferably 110 to 120 ℃. The pre-reaction time may be 4 to 9min, preferably 5 to 8min, more preferably 5 to 7min.
According to one embodiment of the invention, the dried granules are reacted with sulfuric acid at 100-120 ℃ for 5-8 min to obtain a pre-reaction mass. The sulfuric acid is a mixture of 98wt% concentrated sulfuric acid and 105wt% fuming sulfuric acid; the mass ratio of the dried particles to 98wt% concentrated sulfuric acid and 105wt% fuming sulfuric acid is 1 (0.9-1.0) to 0.75-0.8.
< baking step >
And roasting the pre-reaction material at 210-270 ℃ to obtain gas containing hydrogen fluoride and roasting slag. Thus being beneficial to improving the decomposition rate of the rare earth and the niobium and further being beneficial to improving the leaching rate of the rare earth and the niobium. Furthermore, this is advantageous in that a higher yield of hydrogen fluoride gas can be obtained.
The calcination temperature may be 210 to 270 ℃, preferably 220 to 260 ℃, and more preferably 230 to 250 ℃. The calcination time may be 90 to 130min, preferably 100 to 130min, more preferably 100 to 120min, and still more preferably 110 to 120min.
According to one embodiment of the invention, the pre-reaction material is reacted for 90-130 min at 210-270 ℃ to obtain gas containing hydrogen fluoride and roasting slag.
While the calcination is carried out, the generated gas containing the hydrogen fluoride can be washed by a washing tower and then condensed and rectified to obtain the anhydrous hydrogen fluoride. The washed recovered acid can return to the raw material (the dried fluorite concentrate particles) to continue to participate in the reaction, so that the cyclic utilization of the acid is facilitated.
The roasting slag is solid containing rare earth elements, niobium elements and calcium sulfate.
< Leaching step >
Leaching the roasting slag with water, and filtering to obtain leaching solution containing rare earth and niobium and water leaching slag.
The mass ratio of the roasting slag to the water is 1 (0.11-0.2), preferably 1 (0.14-0.2), and more preferably 1 (0.15-0.17). Thus being beneficial to improving the leaching rate of the rare earth and the niobium and improving the purity of the rare earth and the niobium. In the present invention, the mass of the roasting slag is calculated as the dry weight of the roasting slag. According to research and experiments, the invention discovers that the mass ratio of the roasting slag to the water needs to be controlled within the range to obtain higher leaching rates of the rare earth and the niobium, and the invention is favorable for maintaining higher effective contents of leaching solutions of the rare earth and the niobium.
In the invention, an intelligent high-efficiency filter press can be adopted for leaching and filtering to obtain leachate containing rare earth and niobium and water leaching slag. The leaching rate of rare earth is 95% or more, preferably 96% or more, and more preferably 98% or more. The leaching rate of niobium is 85% or more, preferably 86% or more, and more preferably 87% or more. The P content in the leaching solution containing rare earth and niobium is less than 0.045wt%.
The leaching rate of the rare earth and niobium is calculated according to the following formula:
the leaching rate of the rare earth is =100% - [ the mass of the water leaching residue x the REO content in the water leaching residue/(the mass of the fluorite concentrate x the REO content in the fluorite concentrate + the mass of the iron tailings x the REO content in the iron tailings) ]. Times.100%.
Leaching rate of niobium =100% - [ mass of water leaching residue × Nb in water leaching residue 2 O 5 Content/(fluorite concentrate quality x Nb in fluorite concentrate) 2 O 5 Content + quality of iron tailings x Nb in iron tailings 2 O 5 Content)]×100％。
< magnetic separation step >
And carrying out magnetic separation on the obtained water leaching slag by adopting a magnetic separator to obtain calcium sulfate. The magnetic separation intensity is 3.0 to 4.2T, preferably 3.0 to 4.0T, and more preferably 3.5 to 3.9T. Thus, gypsum meeting practical requirements can be obtained. The purity of the obtained calcium sulfate is more than or equal to 90wt%.
< neutralization step >
And (3) neutralizing and removing impurities from the leaching solution containing the rare earth and the niobium by adopting magnesium oxide to obtain the neutralized leaching solution containing the rare earth and the niobium and neutralized slag. The neutralization step may be those known in the art and will not be described herein. The neutralized slag may be returned to the water leaching step for further leaching to recover residual rare earth and niobium.
< analytical method >
The analytical test methods used in the examples and comparative examples are described below:
CaF 2 the content is as follows: the EDTA (ethylene diamine tetraacetic acid) volumetric method is adopted for testing, and the standard GB5195.1-85 is adopted.
REO content: the test is carried out by adopting a gravimetric method according to the standard GB/T6730.25-2021.
Nb 2 O 5 The contents are as follows: the test is carried out by a gravimetric method according to the standard GB/T3654.1-1983.
The content of P: spectrophotometry is adopted according to the standard GB/T5009.87-2003.
TFe: adopting a potassium dichromate titration method according to the national standard GB/T6370.66.
Purity of gypsum: the test was carried out by the barium sulfate precipitation method, according to standard GB/T5484-2012.
Example 1
CaF in the Fluorite concentrate used in this example 2 83.16wt%, REO 6.05wt%, nb 2 O 5 The content was 0.18wt%, and the P content was 0.96wt%. In the used iron tailings, the TFe content is 35.26wt%, and the CaF content 2 19.64wt% of REO, 5.71wt% of Nb 2 O 5 The content was 0.17wt%, and the P content was 1.33wt%.
Mixing fluorite concentrate and iron tailings according to the mass ratio of 1:0.25, preparing balls with the diameter of 2-5 mm on a disc granulator, and drying the prepared balls at 100 ℃ for 24h to obtain dried particles.
And (2) uniformly mixing the dried particles with 98wt% concentrated sulfuric acid and 105wt% fuming sulfuric acid according to a mass ratio of 1.
And leaching and filtering the roasting slag by water by adopting an intelligent high-efficiency filter press in one step to obtain leaching solution containing rare earth and niobium and water leaching slag. And during leaching, the mass ratio of the roasting slag to water is 1.
And (3) neutralizing the leachate containing the rare earth and the niobium by adopting magnesium oxide to obtain the neutralized leachate containing the rare earth and the niobium. The amount of magnesium oxide is 6.0wt% of the fluorite concentrate.
Removing impurities such as iron and thorium from the water leaching slag by a superconducting strong magnetic separator under the condition that the magnetic field intensity is 3.7T, and obtaining high-purity gypsum (namely calcium sulfate). The results are shown in Table 1.
Example 2
CaF in the Fluorite concentrate used in this example 2 85.58wt% of REO, 5.94wt% of Nb 2 O 5 The content was 0.21wt%, and the P content was 0.88wt%. In the used iron tailings, the TFe content is 37.16wt%, and the CaF content 2 27.44wt%, REO 7.31wt%, nb 2 O 5 The content was 0.15wt%, and the P content was 1.19wt%.
Mixing fluorite concentrate and iron tailings according to the mass ratio of 1:0.22, preparing balls with the diameter of 2-5 mm on a disc granulator, and drying the prepared balls at 90 ℃ for 30h to obtain dried particles.
Uniformly mixing the dried particles with 98wt% of concentrated sulfuric acid and 105wt% of fuming sulfuric acid according to the mass ratio of 1.
And leaching and filtering the roasting slag by water by adopting an intelligent high-efficiency filter press in one step to obtain leaching solution containing rare earth and niobium and water leaching slag. When the slag is discharged, the mass ratio of the roasting slag to the water is 1.
And (3) neutralizing the leachate containing the rare earth and the niobium by adopting magnesium oxide to obtain the neutralized leachate containing the rare earth and the niobium. The amount of magnesium oxide is 7.9wt% of the fluorite concentrate.
Removing impurities such as iron and thorium contained in the water leaching slag by a superconducting strong magnetic separator under the condition that the magnetic field intensity is 3.0T, and obtaining high-purity gypsum. The results are shown in Table 1.
Example 3
CaF in the Fluorite concentrate used in this example 2 86.88wt% of REO, 5.85wt% of Nb 2 O 5 The content was 0.14wt%, and the P content was 0.91wt%. Among the iron tailings used, TFe content 37.72wt%, caF 2 22.36wt% of REO, 7.15wt% of Nb 2 O 5 The content was 0.21wt%, and the P content was 1.14wt%.
Mixing fluorite concentrate and iron tailings according to the mass ratio of 1:0.24, preparing balls with the diameter of 2-5 mm on a disc granulator, and drying the prepared balls at 100 ℃ for 24h to obtain dried particles.
Uniformly mixing the dried particles with 98wt% concentrated sulfuric acid and 105wt% fuming sulfuric acid according to the mass ratio of 1.
And leaching and filtering the roasting slag by water by adopting an intelligent high-efficiency filter press in one step to obtain leaching solution containing rare earth and niobium and water leaching slag. Wherein, during leaching, the mass ratio of the roasting slag to water is 1.
And (3) neutralizing the leachate containing the rare earth and the niobium by adopting magnesium oxide to obtain the neutralized leachate containing the rare earth and the niobium. Wherein the amount of the magnesium oxide is 8.2wt% of the fluorite concentrate.
Removing impurities such as iron and thorium contained in the water leaching slag by a superconducting strong magnetic separator under the condition that the magnetic field intensity is 4.0T, and obtaining high-purity gypsum. The results are shown in Table 1.
Example 4
CaF in the Fluorite concentrate used in this example 2 90.75wt%, REO 4.83wt%, nb 2 O 5 The content was 0.22wt%, and the P content was 0.76wt%. In the used iron tailings, the TFe content is 39.16wt%, and the CaF content 2 22.36wt% of REO, 5.47wt% of Nb 2 O 5 The content was 0.16wt%, and the P content was 1.08wt%.
Mixing fluorite concentrate and iron tailings according to a mass ratio of 1:0.15, preparing balls with the diameter of 2-5 mm on a disc granulator, and drying the prepared balls at 100 ℃ for 24 hours to obtain dried particles.
Uniformly mixing the dried particles with 98wt% concentrated sulfuric acid and 105wt% fuming sulfuric acid according to the mass ratio of 1.
And leaching and filtering the roasting slag by water by adopting an intelligent high-efficiency filter press in one step to obtain leaching solution containing rare earth and niobium and water leaching slag. Wherein, during leaching, the mass ratio of the roasting slag to water is 1.
And (3) neutralizing the leachate containing the rare earth and the niobium by adopting magnesium oxide to obtain the neutralized leachate containing the rare earth and the niobium. Wherein the amount of the magnesium oxide is 5.9wt% of the fluorite concentrate.
Removing impurities such as iron and thorium contained in the water leaching slag by a superconducting strong magnetic separator under the condition that the magnetic field intensity is 3.9T, and obtaining high-purity gypsum. The results are shown in Table 1.
Comparative example 1
The same as example 2, except for the following parameter settings:
no iron tailings were added to this comparative example. Wherein the mass ratio of the fluorite concentrate to the 98% concentrated sulfuric acid to the 105% fuming sulfuric acid is 1.
Comparative example 2
The same as example 2, except for the following parameter settings:
no fluorite concentrate was added to this comparative example. Wherein the mass ratio of the iron tailings, 98% concentrated sulfuric acid and 105% fuming sulfuric acid is 1.
Comparative example 3
The difference from the example 2 is that the mass ratio of the dried particles to 98wt% concentrated sulfuric acid and 105wt% fuming sulfuric acid is 1.
TABLE 1
| Numbering | Leaching rate of rare earth% | Leaching rate of niobium% | Purity of gypsum% |
| Example 1 | 96.4 | 89.3 | 91.5 |
| Example 2 | 98.2 | 87.1 | 90.3 |
| Example 3 | 95.9 | 85.6 | 94.2 |
| Example 4 | 95.7 | 86.2 | 92.1 |
| Comparative example 1 | 79.6 | 83.9 | 90.1 |
| Comparative example 2 | 92.4 | 85.7 | 60.5 |
| Comparative example 3 | 91.8 | 84.4 | 90.7 |
The present invention is not limited to the above-described embodiments, and any variations, modifications, and substitutions which may occur to those skilled in the art may be made without departing from the spirit of the invention.
Claims (10)
1. A method for the cooperative treatment of fluorite concentrate and iron tailings is characterized by comprising the following steps:
1) Mixing fluorite concentrate and iron tailings in a mass ratio of 1 (0.1-0.4), granulating, and drying to obtain dried particles; wherein CaF in the fluorite concentrate 2 More than 75wt%, REO 4.5-9 wt%, nb 2 O 5 The content is 0.07 to 0.5 weight percent, and the content of P is 0.5 to 1.4 weight percent; the iron tailings contain 30-45 wt% of TFe, 4.6-8 wt% of REO and Nb 2 O 5 The content is 0.07-0.5 wt%, and the P content is 0.8-1.7 wt%;
2) Pre-reacting the dried particles with sulfuric acid at 100-145 ℃, and then roasting at 210-270 ℃ to obtain gas containing hydrogen fluoride and roasting slag;
3) Leaching the roasting slag with water, and filtering to obtain leaching solution containing rare earth and niobium and water leaching slag.
2. The method of claim 1, wherein the dried particles are processed by steps comprising:
mixing the fluorite concentrate with the iron tailings, preparing the mixture into balls, and drying the prepared balls for 20-38 hours at the temperature of 90-120 ℃ to obtain dried particles.
3. The process of claim 1, wherein in step 2), the sulfuric acid is a mixture of 98wt% concentrated sulfuric acid and 105wt% fuming sulfuric acid.
4. The method of claim 3, wherein the mass ratio of the dried particles, 98wt% concentrated sulfuric acid and 105wt% fuming sulfuric acid is 1 (0.85-1.0): (0.5-0.8).
5. The method according to claim 1, wherein the pre-reaction temperature in step 2) is 100 to 135 ℃ and the pre-reaction time is 4 to 9min.
6. The method as claimed in claim 1, wherein the calcination temperature in step 2) is 210-260 ℃ and the calcination time is 90-130 min.
7. The method of claim 1, wherein in the step 2), the roasting temperature is 220 to 260 ℃ and the roasting time is 100 to 130min.
8. The method according to claim 1, wherein in the step 3), the mass ratio of the roasting slag to the water is 1 (0.11-0.2).
9. The method according to any one of claims 1 to 8, further comprising the steps of:
magnetically separating the water leaching slag obtained in the step 3) by using a magnetic separator to obtain calcium sulfate; wherein, the magnetic separation intensity is 3.0-4.2T.
10. The method of claim 1, further comprising the steps of:
neutralizing the leaching solution containing the rare earth and the niobium obtained in the step 3) by adopting magnesium oxide to obtain neutralized leaching solution containing the rare earth and the niobium and neutralized slag.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN202211522585.4A CN115821078A (en) | 2022-11-30 | 2022-11-30 | Method for cooperatively treating fluorite concentrate and iron tailings |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CN202211522585.4A CN115821078A (en) | 2022-11-30 | 2022-11-30 | Method for cooperatively treating fluorite concentrate and iron tailings |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| CN115821078A true CN115821078A (en) | 2023-03-21 |
Family
ID=85533201
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CN202211522585.4A Pending CN115821078A (en) | 2022-11-30 | 2022-11-30 | Method for cooperatively treating fluorite concentrate and iron tailings |
Country Status (1)
| Country | Link |
|---|---|
| CN (1) | CN115821078A (en) |
Citations (8)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN101492771A (en) * | 2009-03-04 | 2009-07-29 | 广州有色金属研究院 | Hingganite ore concentrate decomposition method |
| CN102703697A (en) * | 2012-06-29 | 2012-10-03 | 广州有色金属研究院 | Method for recovering rare earth-niobium-ferrum paragenic ore |
| CN103415631A (en) * | 2011-01-06 | 2013-11-27 | 阿海珐矿业公司 | Dissolution and recovery of at least one element nb or ta and of at least one other element u or rare earth elements from ores and concentrates |
| CN106636614A (en) * | 2017-01-17 | 2017-05-10 | 东北大学 | Method for leaching niobium, scandium and rare earth elements from tailings |
| CN106636691A (en) * | 2016-12-28 | 2017-05-10 | 核工业北京化工冶金研究院 | Method used for extracting uranium and niobium from low-grade ore |
| CN106916941A (en) * | 2017-03-13 | 2017-07-04 | 北京科技大学 | A kind of method that utilization iron content rare-earth original ore separates production rare earth |
| CN109371239A (en) * | 2018-12-03 | 2019-02-22 | 包头稀土研究院 | A method of low-grade fluorite mine of the processing containing rare earth |
| CN109837400A (en) * | 2018-11-14 | 2019-06-04 | 中国地质科学院郑州矿产综合利用研究所 | Method for extracting rare earth and niobium from niobium-rich slag |
-
2022
- 2022-11-30 CN CN202211522585.4A patent/CN115821078A/en active Pending
Patent Citations (8)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN101492771A (en) * | 2009-03-04 | 2009-07-29 | 广州有色金属研究院 | Hingganite ore concentrate decomposition method |
| CN103415631A (en) * | 2011-01-06 | 2013-11-27 | 阿海珐矿业公司 | Dissolution and recovery of at least one element nb or ta and of at least one other element u or rare earth elements from ores and concentrates |
| CN102703697A (en) * | 2012-06-29 | 2012-10-03 | 广州有色金属研究院 | Method for recovering rare earth-niobium-ferrum paragenic ore |
| CN106636691A (en) * | 2016-12-28 | 2017-05-10 | 核工业北京化工冶金研究院 | Method used for extracting uranium and niobium from low-grade ore |
| CN106636614A (en) * | 2017-01-17 | 2017-05-10 | 东北大学 | Method for leaching niobium, scandium and rare earth elements from tailings |
| CN106916941A (en) * | 2017-03-13 | 2017-07-04 | 北京科技大学 | A kind of method that utilization iron content rare-earth original ore separates production rare earth |
| CN109837400A (en) * | 2018-11-14 | 2019-06-04 | 中国地质科学院郑州矿产综合利用研究所 | Method for extracting rare earth and niobium from niobium-rich slag |
| CN109371239A (en) * | 2018-12-03 | 2019-02-22 | 包头稀土研究院 | A method of low-grade fluorite mine of the processing containing rare earth |
Similar Documents
| Publication | Publication Date | Title |
|---|---|---|
| CN107245573B (en) | A method of synthetical recovery zinc, germanium, lead, silver from Second-rate zinc oxide powder | |
| CN106048217B (en) | The comprehensive reutilization method of oxide powder and zinc | |
| CN105087905A (en) | Method for treating transition type and limonite type nickel laterite ore by using industrial pickling waste acid | |
| CN107099672B (en) | The recovery method of the fume from steel making containing zinc | |
| CN112239808B (en) | Method for extracting vanadium from vanadium-containing high-calcium high-phosphorus slag | |
| CN103667695A (en) | Method for extracting arsenic from gold ore | |
| CN106882839B (en) | Method for comprehensively utilizing titanium white waste acid | |
| CN115821078A (en) | Method for cooperatively treating fluorite concentrate and iron tailings | |
| CN108163880B (en) | Method for preparing gypsum powder by using zinc smelting waste acid | |
| CN111100996A (en) | Method for preparing vanadium oxide from acidic low-concentration vanadium liquid | |
| CN103613116B (en) | Method for simultaneously and comprehensively recycling zinc ash and high-arsenic zinc oxide in iron and steel plant | |
| CN109943722A (en) | A kind of reuse method of secondary zinc oxide | |
| CN104789784A (en) | Zinc pyrometallurgical smoke-dust recovery processing process | |
| CN112391537B (en) | Method for extracting vanadium by using hydrochloric acid, sulfuric acid and vanadium-containing high-calcium high-phosphorus slag | |
| CN109796049B (en) | Method for preparing iron oxide red by using iron slag precipitated by zinc hydrometallurgy goethite method | |
| CN114192274A (en) | Manganese ore resource utilization method combining smelting and selecting materials | |
| CN115852174A (en) | Method for recovering rare earth and niobium from fluorite concentrate | |
| CN115992310A (en) | Method for treating fluorite concentrate containing monazite | |
| CN111017974A (en) | Mineral processing technology for resource utilization of low-grade celestite | |
| CN108913887A (en) | A method of zinc oxide is produced using containing zinc ore crude | |
| CN110972482B (en) | Beneficiation method for low-grade zinc-containing raw ore | |
| CN116798760B (en) | Method and flux for preparing soft magnetic manganese zinc ferrite composite material by utilizing manganese anode slime and zinc slime | |
| CN108866335A (en) | A method of copper and zinc are effectively recycled from the pickle liquor of Gold Concentrate under Normal Pressure | |
| CN114854990B (en) | Method for recovering rare earth from waste CRT fluorescent powder | |
| CN111020201B (en) | Method for purifying and removing impurities from smelting flue gas desulfurization gypsum |
Legal Events
| Date | Code | Title | Description |
|---|---|---|---|
| PB01 | Publication | ||
| PB01 | Publication | ||
| SE01 | Entry into force of request for substantive examination | ||
| SE01 | Entry into force of request for substantive examination |