WO1996012047A1 - Procede de recuperation de titane et de vanadium - Google Patents

Procede de recuperation de titane et de vanadium Download PDF

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Publication number
WO1996012047A1
WO1996012047A1 PCT/GB1995/002454 GB9502454W WO9612047A1 WO 1996012047 A1 WO1996012047 A1 WO 1996012047A1 GB 9502454 W GB9502454 W GB 9502454W WO 9612047 A1 WO9612047 A1 WO 9612047A1
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WIPO (PCT)
Prior art keywords
slag
vanadium
titanium
roasting
mixture
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Application number
PCT/GB1995/002454
Other languages
English (en)
Inventor
Peter Tyson
James Hamilton
Original Assignee
Magmint Limited
Civelli, Carlo, Giuseppe
Priority date (The priority date 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 date listed.)
Filing date
Publication date
Application filed by Magmint Limited, Civelli, Carlo, Giuseppe filed Critical Magmint Limited
Priority to AU36596/95A priority Critical patent/AU3659695A/en
Publication of WO1996012047A1 publication Critical patent/WO1996012047A1/fr

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Classifications

    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • C22B7/04—Working-up slag
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B34/00—Obtaining refractory metals
    • C22B34/10—Obtaining titanium, zirconium or hafnium
    • C22B34/12—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08
    • C22B34/1204—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 preliminary treatment of ores or scrap to eliminate non- titanium constituents, e.g. iron, without attacking the titanium constituent
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B34/00—Obtaining refractory metals
    • C22B34/10—Obtaining titanium, zirconium or hafnium
    • C22B34/12—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08
    • C22B34/1218—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by dry processes
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B34/00—Obtaining refractory metals
    • C22B34/10—Obtaining titanium, zirconium or hafnium
    • C22B34/12—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08
    • C22B34/1236—Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B34/00—Obtaining refractory metals
    • C22B34/20—Obtaining niobium, tantalum or vanadium
    • C22B34/22—Obtaining vanadium
    • 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

Definitions

  • the present invention relates to a titanium and optionally a vanadium recovery process.
  • the metals are extracted from slag material.
  • the present invention also relates to a titanium, vanadium and iron recovery process from vanadiferous titano- agnetite.
  • Titanium and vanadium are industrially important chemical elements. Titanium dioxide (Ti0 2 ) is widely used as a white pigment in paints and inks, a paper filler, a colouring agent for rubber and leather, a component of ceramics and as an opacifying agent in porcelain enamels. Because of its high dielectric constant it has found use in dielectrics. Ti0 2 and titanium metal producers generally prefer to use feedstocks that possess a relatively high titanium content. This has enabled them to economise in their use of sulphuric acid or chlorine in the recovery process and to curtail the extent of pollution arising at the recovery plants. Hence, natural rutile has traditionally been the preferred raw material for most titanium metal and chloride-route Ti0 2 manufacturers. The technical capability for recovering titanium from significantly lower-grade titaniferous magnetites, for example those with a titanium content of less than 30% has to-date been limited.
  • Vanadium metal is by the steel industry as an alloying agent, in particular, in the production of carbon and alloy steels.
  • Vanadium compounds have high utilization in the manufacture of oxidation catalysts for the chemical industry and in the ceramic industry as colouring agents. Vanadium is in many cases recovered as a co-product or by-product of other elements with which it is associated for example iron. Vanadium (and uranium) are extracted from carnotite and similar ores.
  • AU-A-41686/93 describes the production of a titanium rich slag from titano agnetite.
  • the titanomagnetite is fed, together with a carbonaceous reductant, into a fluxless arc furnace.
  • a titanium dioxide product suitable for the sulphate process of pigment production is recovered and pig iron containing vanadium is tapped off as a by product.
  • EP-A-0583126 describes the production of a titanium rich slag and pig iron from ilmenite.
  • the il enite is fed continuously together with a carbonaceous reductant in the absence of fluxes, to an arc furnace.
  • Titanium rich slag which can be used as the feed for chlorine based titanium dioxide production is recovered and pig iron is tapped off (with any impurities) as a by product.
  • One primary vanadium operation (where only the extracted vanadium is given value) from titaniferous magnetite ore involves a roast-leach process using sodium carbonate as the leachant. Valuable vanadium pentoids are produced. The iron and titanium values are discarded as waste products.
  • Extraction technology has enabled both vanadium and titanium to be extracted from ore following upgrading of the ore to a concentrate (with the disposal of a 55% waste fraction) .
  • the vanadium is extracted by a hydrometallurgical method prior to reduction and smelting to separate the Ti0 2 in the slag from the iron.
  • a hydrometallurgical method prior to reduction and smelting to separate the Ti0 2 in the slag from the iron.
  • such a process which extracts the vanadium initially from the ore is highly inefficient and not commercially viable when the ore contains a relatively low vanadium concentration.
  • the present invention provides a process for the recovery of titanium and vanadium from a slag comprising mixing the slag with an alkali metal salt, roasting the mixture in oxygen and adding a leaching liquid to the mixture from which vanadium is recovered.
  • Such a recovery process enables the economic recovery of titanium and vanadium from ores which are relatively poor in these elements, in particular from ores with less than 30% titanium or less than 1.7% vanadium or from ores which because of their composition cannot be upgraded to a concentrate prior to the recovery process.
  • the present invention thus allows for utilisation of many of the world's previously unexploitable vanadiferous titano- agnetites.
  • the present invention also provides a process for the upgrading of Ti slag to remove, amongst others, vanadium impurities, comprising mixing the slag with a alkali metal salt and roasting the mixture in oxygen at a temperature in the range of 900 to 1200°C.
  • Preferred features of the recovery process described herein are also applicable to the upgrading process.
  • the present invention describes a process for the separation of V from Ti from a slag/ore containing both metals.
  • Either metal may be present in minimum quantities or both metals may be present in substantially greater than minimum quantities.
  • any large lumps of the slag Prior to the roasting step, it is preferred that any large lumps of the slag are crushed to reduce their size. Crushing may be achieved by known methods for example, by mechanical jaws. Screening removes oversized lumps which may be recycled and recrushed.
  • any iron prill removed By the terms iron prill are meant any lumps of iron "shot” resulting from a previous smelting process. Such iron prill removed at this stage may be recovered and upgraded to high quality pig iron.
  • the reduced size slag may be further reduced in size by ball milling.
  • the particle size should not exceed 200 ⁇ m. More preferably the particle size should be less than 45 ⁇ m (diameter) . This size reduction increases the surface area of the slag for effective reaction during the recovery process. Removal of "contaminating" iron prills increases the effective concentration of titanium and vanadium in the slag prior to the recovery process.
  • the slag and alkali metal salt are in the ratio of from 3:1 to 1:1 and more preferably approximately 5:1 weight by weight ratio respectively. This ratio results in the highest vanadium and titanium recovery.
  • the mixture of the slag and the alkali metal salt are wetted, pelletized and dried before the roasting step.
  • a rotating drum may be provided to produce pellets of a pre-determined size.
  • the preferred pellet size is one-quarter to three eighths of an inch (diameter) .
  • the damp pellets are dried in a pre-calciner to prevent superheating and splitting during roasting.
  • the roasting of the slag and the alkali metal salt mixture in oxygen is carried out at a temperature below the temperature required to fuse the slag with the alkali metal salt and below the te perative to cause sintering.
  • the process is preferably carried out in the range of from 900 to 1200°, more preferably in the ranges 1000 to 1200°C, 1050 to 1150°. The most preferred temperature is approximately 1125°C.
  • the best extraction results are achieved when the roasting is continued for from 20 to 60 minutes, preferably 25 to 35 inures, more preferably approximately 30 minutes.
  • the roasting process is carried out in a vertical roaster.
  • Alternative known roasting means such as a rotating furnace may be used.
  • the oxygen may be supplied to the mixture during roasting by any known means but preferably as an upward flow.
  • the oxygen should be supplied constantly throughout the process so as to maintain an oxygen rich atmosphere at least during loading and roasting.
  • the mixture After roasting the mixture is preferably cooled and quenched by the addition of a leach liquid at from 85 to 95°, preferably approximately 90°C and left for a period of from 20 to 60 minutes, preferably from 25 to 35 minutes, more preferably 30 minutes. At this stage, the soluble vanadium transfers from the slag to the leach liquid.
  • a leach liquid at from 85 to 95°, preferably approximately 90°C and left for a period of from 20 to 60 minutes, preferably from 25 to 35 minutes, more preferably 30 minutes.
  • the leach liquid may be withdrawn from the mixture at this stage or may be recycled to the next charge.
  • the leach liquid is withdrawn when a pre-determined vanadium concentration is reached.
  • the leach liquid may have a slightly acidic or alkaline pH. Aluminium sulphate is added to the leach liquid which is then adjusted to a pH of from 7 to 9 , preferably 7.5 to 8.5, more preferably approximately 8 by the addition of alkali or acid co-precipitated silica and/or alumina impurities can be filtered off.
  • the leach liquid is preferably evaporated until a vanadium content of 40 to 80 g/1, preferably 65 to 75 g/1 and more preferably approximately 70g
  • the vanadium may then be brought into solid form by the addition of ammonium chloride or ammonium sulphate to form solid ammonium metavanadate.
  • the leach liquid is water or a caustic solution, for example sodium hydroxide or sodium carbonate.
  • the remaining slag is preferably further leached in dilute sulphuric acid or hydrochloric acid, washed and dried to form an upgraded titanium dioxide slag.
  • the alkali metal salt is sodium carbonate, sodium hydroxide, sodium sulphate or a mixture of two or more thereof. More preferably the alkali metal salt is sodium carbonate and sodium sulphate in a ratio of from 2:3 to 3:2, preferably 1:1 by weight. This combination produces high titranium and vanadium recovery levels with low levels of impurities .
  • the present invention also provides a process for the extraction of iron, vanadium and titanium from vanadiferous titano-magnetite comprising pre-reducing the magnetite with a carbon reductant, smelting the ore in the absence of flux with -a- carbon reductant, recovering an iron phase and a slag phase wherein the slag phase comprises substantially all of the titanium and vanadium and recovering the titanium and vanadium by the process as hereinbefore described.
  • the present invention firstly removes the iron phase to leave a slag containing soluble vanadium and titanium.
  • the vanadium is solublized and then extracted by leaching and the titanium bearing slag is upgraded by acid digestion and subsequent washing.
  • Conditions as stated above result in a slag of acceptable V 2 0 5 content ( ⁇ 0.6%) , unattackable Ti0 2 (0.2%) and low Ti 2 o. values ( ⁇ 0.1%) , all necessary for successful TiO. extraction.
  • composition of the slag lends itself to an alkali attack in that the Ti will not be affected at low temperatures (1200°C) , iron will not be attacked but silicon, aluminium and vanadium may be converted to soluble forms.
  • Various additions of sodium carbonate were made to slag which was ground and passed through a 45 micron sieve and furnaced at 1000°C. The following results were obtained as depicted in the graph.
  • pellets of these particle sizes gave satisfactory vanadium recovery.
  • Pellets with 45 micron size particles gave a more intimate mix than larger particles.
  • Pellet sizes of from one-eighth to 1 inch, preferably one- quarter to one-half, more preferably ⁇ -3/8 and 3/8- are used. Analysis throughout has given a steady Vanadium recovery of approximately 80+% and Al-,0 3 of 4%, Si0 2 6% and Ti0 2 ⁇ 1% in all water leach liquors.
  • Acid leach of the original slag showed a 23% removal of FeO and relatively low vanadium. After roasting the FeO removal was low (oxidation state of iron) .
  • Suitable furnace lining can also reduce the MgO.
  • the conventional method of obtaining magnetic concentrate involves fine grinding and separation of magnetic phases containing Ti (Ti- agnetites) . In this process ilmenite and silicates are discarded. Grinding prior to pre-reduction places a rigorous constraint on the pre-reduction mechanics, in that the process must be carried out in a fluidised bed system which is expensive and is generally unknown technology, or a pelletised material in a rotary furnace where choice of binder with low Al 2 0 3 is critical .
  • the feedstock to the process can be obtained in either of two ways: the ore can be selectively mined to extract only the solid magnetite units from the bulk layer (about 7-10 metres of material in 1-3 metre units) .
  • the other method is to bulk mine the deposit, crush to 10-20 mm then use a preliminary magnetic separation stage to discard the silicate-rich material.
  • the size of crushed ore is determined empirically.
  • An improved pre-reduction route makes use of existing technology known in the art.
  • Coarsely- crushed material is introduced into a rotary kiln with coal as a reductant.
  • the furnace is fired by pulverised fuel (PF) , which is relatively cheap per heat unit.
  • PF pulverised fuel
  • the coarse size (preferably 10-20mm) of the feed prevents choking the furnace and unwanted clinkering, and some fines (perhaps 10-20%) can be accommodated.
  • Such a process is able to yield 70% pre-reduction on 12% Ti0 2 magnetite. Similar efficiencies on 21% Ti0 2 feedstock are about 60% pre- reduced.
  • the material On exiting the furnace, the material is quenched or air-cooled, and introduced into a ball/rod mill for comminution to the experimentally-determined size grading. Magnetic separation is now carried out. This step allows the previously relatively-nonmagnetic ilmenite to report to the magnetic fraction and all gangue is discarded.
  • the concentrate will have Ti0 2 values of about 27-29% with SiO- of ⁇ 1% and A1 2 0 3 in the 1.5% range.
  • FeO was higher in the slag phase, with Ti0 2 contents between 66-75%. This slag also contained some lower oxides of titanium.
  • the process as described in Claim 1 may be applied to either slag, either roasting under oxidising conditions to reform TiO-, followed by leaching (including hot leaching).
  • Ti0 2 content can be increased by this process from 85 to 92%, whilst Al 2 0 j , Si0 2 and V 2 0 5 would show corresponding decreases.
  • the process may be described as follows: the raw (not pre-reduced) concentrate is smeltered on a 200KVA DC plasma furnace under a range of conditions. Over-reducing conditions brought about by high proportions of reductant (supra- stoichiometric) yielded metal with high levels of V and TiC.
  • the slags have native iron, TiC and Al 2 0 3 , MgO and CaO, with 85% Ti0 2 or better, mostly in the form of lower Ti-oxides.
  • the furnace conditions With a decrease in the amount of reductant, the furnace conditions become relatively less reducing and V exhibits its characteristic switch of redox-driven partitioning into the slag phase.
  • the metal is free of TiC and V, yet contains the same proportions of Mn, Cr and most traces to the over-reduced metal.
  • the system can be run at conditions of over-reduction.
  • This does have some drawbacks.
  • the five main ones are: 1) higher liquidus temperatures; 2) the A1-.0-. content of the system is raised because of the larger amounts of coal used; 3) some Ti reports to the metal in the carbide phase; 4) slag volumes are smaller; 5) lower Ti-oxides are prevalent.
  • the fact that V reports to the metal is may be advantaeous, as may be the high Ti0 2 grade of the slag and the salvaging of Fe from this slag.
  • the upgrading of the Ti rich slag, to remove, in particular vanadium impurities is a process described herein according to the present invention
  • point 2 above is negated.
  • point 1 is bypassed.
  • Points 3 and 4 can be dealt with using technology known in the art.
  • the iron from the furnace may be tapped into a shaking ladle, which is then agitated and soft-blown with oxygen to release the V and to oxidise the TiC (processes known in the art) . Both are gathered into a slag which can be mechanically skimmed. This slag can be used in the production of Fe and V.
  • the metal can be treated with FeSi and desulphurised as required.
  • the over-reduced nature of the metal allows easier adherence to the strict C- content requirements of the pig-users.
  • the conditions of over-reduction are 1) a reducing agent such as low- ash coal and 2)heat in the range of 1400 to 1900°C.
  • the roasting step in the presence of an alkali metal salt (as claimed in claim 1) is carried out on a slag rich in Ti, but not rich in 2 0 5 . Any vanadium present in the slag is in minimum quantities (the majority having been extracted via the shaking ladle step) .
  • the slag produced from the relatively oxidising conditions of the route contains about 66-75% Ti0 2 , some as lower oxides.
  • the rest of the slag is made up of FeO (approximately 8-15%) with abundant A1 2 0 3 (approximately 5%) , Si0 2 CaO, MgO and, of course, V 2 0 5 .
  • the process as described hereinbefore according to the invention upgrades the slag into a more marketable form.
  • leaching with H 2 S0 4 at a temperature in the range of from 40 to 90°C removes up to 70% of the remaining Al 2 0 j and Si0 2 .
  • the preferred temperature of the hot leach step is 60 to 85°C, more preferably around 80°C.
  • the slag is heated in an oxidising furnace, then two advantages are identified. Firstly, the low Ti oxides are mostly converted to Ti0 2 , which is highly soluble in the sulphate route. Hot leaching on this material then removes the V 2 0 5 and the bulk of the deleterious elements, to yield an acceptable slag for either the sulphate-route producers or for chlorinatable-route producers. Further work on the process is possible in order to remove the bulk of the FeO, thus making a synrutile-like product.
  • the "hot leaching'' may be carried out by adding hot H 2 SO A to cool pellets, by adding cool H 2 S0 4 to the hot pellets, or even by heating a combination of cool pellets with cool H 2 SO .
  • a water leach step (previously used to leach the V) is not required.
  • the high-temperature (up to and around 85%) Ti0 2 slag from the furnace when run at reducing conditions may be granulated by quenching.
  • the quenching serves two purposes. Firstly, it reduces the size of the material without grinding and, secondly, some oxidation takes place by reaction with aerated waters.
  • the process as described in claim 1 may then be applied with roasting in oxiding conditions to reform the lower Ti oxides, followed by leaching (including hot leaching) .
  • This leaching process will remove the bulk of the A1 2 0 3 and the final content of the slag will be around 2-3% following smelting of pre-reduced materal with char. Because the highly-reduced slag has a low FeO content, the amount of Fe to be removed will consequently be less, and .will decrease the volume of circulating leachates. Likewise for the vanadium there may be very little, because it will have reported to the metal. However, the leach process provides a valuable clean-up of the slag. A composition of 92-94% Ti0 2 is possible by optimization of leach conditions. - Final products
  • the trace elements in the metal will be controlled by those siderophile elements originally in the magnetite concentrate, like Mn and Cr which may be at low concentrations in the original ore.
  • the content of phoshorus will be low, because all apatite will have been removed in the magnetic separation of the feedstock.
  • the slag produced according to the invention will be close to ⁇ ynrutile in composition, with approximately 92-94% Ti0 2 as higher oxides, 2-3% A1 2 0 3 , ⁇ 1% Si0 2 CaO, MgO and C, with FeO providing the balance.
  • the content of Cr is low and most V will have leached out.
  • the radioactive elements like Th and U will be approximately in the 3-8 ppm range.
  • the final makeup of the vandium-bearing products will depend on the process followed.
  • the Ti-V slag from the shaking ladle can be readily transformed into V 2 0 5 , and that oxide is also available from the final clean-up of the slag. It has been proposed that V 2 0 5 , Fe and Al can be combined by exothermic reaction into FeV.
  • Fig. 1 is a schematic diagram of a recovery process of vanadium and titanium from slag
  • Fig. 2 is a schematic diagram of a second recovery process of vanadium and titanium from slag.
  • Fig. 3 is a schematic diagram of a recovery process of iron, vanadium and titanium from vanadiferous titano-magnetite.
  • Fig.4 is a schematic diagram of a second recovery process of iron, vanadium and titanium from vanadiferous titano-magnetite including mining and milling stages and a plasma furnace stage.
  • the slag stock pile material is obtained from ore smelting. Large lumps of the slag stock pile are fed into mechanical jaws to reduce size. Screening removes oversized lumps which are recycled, re-crushed and re-screened. Contaminating iron prill is removed at this stage.
  • the reduced-size slag fine slag
  • Fine slag and sodium carbonate are mixed (100 tonnes slag with 20 tonnes sodium carbonate) before wetting and pelletizing.
  • a rotating drum produces pellets of the pre-determined to 1/8 inch (diameter) size.
  • the damp pellets are dried in a pre-calciner to prevent superheating and splitting.
  • the pellets are poured into a vertical roaster. Oxygen flows upwards through the vertical roaster. The charge is roasted at 1125°C for 30 minutes. The hot roasted pellets are removed from the roaster, allowed to cool from red hot and quenched in water at 90°C. They are then leached for 30 mins. The leach liquor is recycled to the next charge until a predetermined vanadium concentration is met.
  • the liquor is treated at a desilication plant to remove silica and alumina. The liquor passes through an evaporator and is evaporated to approximately 70g/litre of vanadium. Ammonium chloride is added to the liquor for ammonium metavanadate precipitation. The remaining slag pellets are further leached in dilute sulphuric acid, washed with water, dried at 110°C and packed for transport.
  • Fig. 2 This process is similar to the process as shown in Fig. 1.
  • the slag stock pile is treated in the same manner until after balled milling and storage in the fine slag bunker.
  • the fine slag is then digested with 2.5% sulphuric acid for 5 hours.
  • the resulting slurry is centrifuged.
  • the resulting liquor is recycled to exhaustion, neutralized with lime and then dumped.
  • the damp slag from centrifugation is mixed with sodium carbonate and sodium sulphate (100 tonnes: 10 tonnes: 10 tonnes respectively) .
  • the damp pellets are then treated as described in Fig. 1, except that the water leached pellets of titanium bearing slag are leached with 2.5% sulphuric acid at 80°C for 30 minutes.
  • the leach liquor is recycled to exhaustion, neutralized with lime and then dumped.
  • the pellets are finally dried at 110°C and packed for transportation.
  • Vanadiferous titano-magnetite is mined from the ground. It then undergoes beneficiation including crushing, grinding and magnetic separation to maximise the concentration of economic materials in the magnetite. The ore then undergoes pre-reduction to increase the ratio of titanium to iron. The resulting mass then undergoes fluxless smelting in an electric arc furnace with carbon reductant. The ratio of carbon to magnetite concentration is precisely determined so as to leave approximately 10% unreduced FeO in the slag. This ensures that the vanadium reports to the slag phase and that the metal phase is as pure in iron as possible. Pig iron is tapped from the lower of two tap holes in the furnace. From this pig iron high quality ductile Iron pigs are manufactured.
  • Slag trapped from the upper of the two tap holes in the furnace is subjected to a titanium and vanadium recovery process as hereinbefore described (beneficiation) .
  • remediation is meant any step or steps in the concentration and/or further processing of an ore (either metallic or non-metallic) .
  • the results are saleable products of titanium dioxide slag and ammonium metavanadate.
  • vanadiferous titano-magnetite is mined from the ground. This may be done selectively as previously described. The mined deposit is crushed to 10-20mm and undergoes a beneficiation circuit to obtain a coarse primary product. Coarse waste is rejected. The coarsely crushed material is then introduced into a rotary kiln with pulverized fuel coal as the reductant. On exiting the furnace the material is quenched with air or water. Off gases are re-routed to pre-heat the furnace feed. The material is then milled to size of less than 75 ⁇ m. Magnetic separation is now carried out to isolate the magnetic magnetite, Ti-variants and magnetic ilmenite.
  • Non-magnetic silicates, phosphates and ash are rejected.
  • the material is then fed into a plasma furnace.
  • the furnace is run a reducing conditions with low-ash coal as the reducing agent.
  • the furnace is operated at the preferred temperature of approxaminately 1700°C.
  • the majority of the vandium reports to the metal phase.
  • the slag includes minimum vanadium and approximately 86% Ti0 2 as some lower oxides. Off gas from the stage may be re-routed to feed pre-heaters.
  • the metal phase (iron) is tapped into a shaking ladle according to known technology. This is agitated and soft blown with oxygen to release a vanadium and Ti0 2 containing slag which may be transferred to a ferrovanadium plant if required.
  • the Si and S content of the iron is adjusted by conventional processes. From this, pig iron is manufactured.
  • the slag (minimum V and about 86° Ti0 2 ) meanwhile is granulated by quenching.
  • the slag then undergoes a process as described hereinbefore, according to the invention, with roasting in oxidising conditions followed by leaching (preferably hot leaching) .
  • This process upgrades the slag by extracting the remaining V 2 0 5 (which may be precipitated) , and removing the majority of Si0 2 , A1 2 0 3 , CaO and MgO impurities.
  • the final slag composition comprises approximately 92 to 94% Ti0 2 .
  • the present invention can also be used to recover either titanium or vanadium from slag or to recover any one or a selection of two from iron, titanium or vanadium from vanadiferous titano-magnetite.

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Abstract

Selon un procédé de récupération ou de séparation de titane et de vanadium contenus dans des scories, on mélange les scories avec un sel de métal alcalin et on calcine le mélange dans de l'oxygène.
PCT/GB1995/002454 1994-10-17 1995-10-17 Procede de recuperation de titane et de vanadium WO1996012047A1 (fr)

Priority Applications (1)

Application Number Priority Date Filing Date Title
AU36596/95A AU3659695A (en) 1994-10-17 1995-10-17 Titanium and vanadium recovery process

Applications Claiming Priority (2)

Application Number Priority Date Filing Date Title
GB9420937A GB2294255A (en) 1994-10-17 1994-10-17 Vanadium recovery process
GB9420937.6 1994-10-17

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CN102061397A (zh) * 2010-06-02 2011-05-18 四川龙蟒矿冶有限责任公司 一种从钒钛磁铁矿中回收利用钒、铬、钛、铁的方法
CN103757199A (zh) * 2013-12-05 2014-04-30 中国科学院过程工程研究所 一种利用高铬型钒钛磁铁精矿制备钒铬钛渣的方法
CN103962220A (zh) * 2014-04-23 2014-08-06 鞍钢集团矿业公司 利用碱浸、酸洗、脱泥及重磁联合再选钒钛磁铁精矿方法
CN106065435A (zh) * 2016-07-18 2016-11-02 江苏省冶金设计院有限公司 一种处理钒渣的方法与系统
CN110699554A (zh) * 2019-10-16 2020-01-17 中冶赛迪工程技术股份有限公司 一种富钒渣生产富钒铁的方法
CN114672645A (zh) * 2022-03-30 2022-06-28 攀枝花学院 利用钒钛磁铁矿尾矿制备钛铁合金的方法

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AR007955A1 (es) * 1996-07-24 1999-11-24 Holderbank Financ Glarus Procedimiento para separar titanio y/o vanadio de hierro bruto
CN1074050C (zh) * 1999-06-12 2001-10-31 湖南省安化县碳化硅厂 一种从含钒石煤中提钒的焙烧方法
CN100357470C (zh) * 2005-09-27 2007-12-26 梅卫东 用钒钛铁精矿制取钛铁、钢及钒铁的方法
DK2572006T3 (da) * 2010-05-19 2019-11-04 Tng Ltd Fremgangsmåde til ekstraktion og indvinding af vanadium
CN101914716B (zh) * 2010-09-03 2011-08-31 四川省达州钢铁集团有限责任公司 钢的熔炼方法及钢筋的生产方法
NZ712526A (en) * 2013-05-17 2017-03-31 Inst Process Eng Cas Method for processing vanadium-titanium magnetite finished ores by using wet process
CN103572062A (zh) * 2013-10-17 2014-02-12 攀钢集团攀枝花钢铁研究院有限公司 一种回收除硅渣中钒的方法
WO2015081775A1 (fr) * 2013-12-05 2015-06-11 中国科学院过程工程研究所 Procédé pour utiliser complètement un concentré de magnétite de vanadium-titane à teneur élevée en chrome
CN111363926B (zh) * 2020-04-26 2021-11-23 攀钢集团钒钛资源股份有限公司 钒渣浅氧化焙烧分离钒的方法

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GB848230A (en) * 1956-08-27 1960-09-14 Blancs De Zinc De La Mediterra A process of separately recovering aluminium, iron and titanium values from material such as "red mud" containing said values
GB1026691A (en) * 1961-04-21 1966-04-20 Laporte Titanium Ltd Improvements in or relating to the treatment of titanium bearing ores
US3816589A (en) * 1971-03-15 1974-06-11 Union Carbide Corp Process for recovery of vanadium values from ferrophosphorus and/or ferrophosphorus mixture
US3929460A (en) * 1973-06-25 1975-12-30 Billiton Research Bv Process for the preparation of vanadium, vanadium alloys or vanadium compounds
DE2656683A1 (de) * 1976-12-15 1978-06-29 Vaw Ver Aluminium Werke Ag Verfahren zur herstellung von ammoniumvanadat
DE2732854A1 (de) * 1977-07-21 1979-02-01 Elektrometallurgie Gmbh Verfahren zum aufbereiten chromhaltiger vanadinschlacken
US4645651A (en) * 1984-01-25 1987-02-24 GFE Geselschaft fur Elektrometallurgie mbH Method of producing vanadium compounds from vanadium-containing residues
US4748009A (en) * 1985-10-05 1988-05-31 Gfe Gesellschaft Fur Elektrometallurgie Mbh Method of recovering vanadium from vanadium-containing materials with at least 6 wt % oxidic vanadium compounds
GB2194941A (en) * 1986-09-10 1988-03-23 Uralsky Inst Chernykh Metall Process for recovering vanadium values
DE3711371A1 (de) * 1987-04-04 1988-10-20 Metallgesellschaft Ag Verfahren zur gewinnung von v(pfeil abwaerts)2(pfeil abwaerts)o(pfeil abwaerts)5(pfeil abwaerts)
AU656476B2 (en) * 1992-07-03 1995-02-02 Mintek The recovery of titanium from titanomagnetite
EP0583126A1 (fr) * 1992-08-11 1994-02-16 Mintek Production de scanès titanifères à partir d'ilménite

Cited By (9)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN102061397A (zh) * 2010-06-02 2011-05-18 四川龙蟒矿冶有限责任公司 一种从钒钛磁铁矿中回收利用钒、铬、钛、铁的方法
CN103757199A (zh) * 2013-12-05 2014-04-30 中国科学院过程工程研究所 一种利用高铬型钒钛磁铁精矿制备钒铬钛渣的方法
CN103757199B (zh) * 2013-12-05 2016-02-17 中国科学院过程工程研究所 一种利用高铬型钒钛磁铁精矿制备钒铬钛渣的方法
CN103962220A (zh) * 2014-04-23 2014-08-06 鞍钢集团矿业公司 利用碱浸、酸洗、脱泥及重磁联合再选钒钛磁铁精矿方法
CN103962220B (zh) * 2014-04-23 2016-02-03 鞍钢集团矿业公司 利用碱浸、酸洗、脱泥及重磁联合再选钒钛磁铁精矿方法
CN106065435A (zh) * 2016-07-18 2016-11-02 江苏省冶金设计院有限公司 一种处理钒渣的方法与系统
CN110699554A (zh) * 2019-10-16 2020-01-17 中冶赛迪工程技术股份有限公司 一种富钒渣生产富钒铁的方法
CN114672645A (zh) * 2022-03-30 2022-06-28 攀枝花学院 利用钒钛磁铁矿尾矿制备钛铁合金的方法
CN114672645B (zh) * 2022-03-30 2024-01-30 攀枝花学院 利用钒钛磁铁矿尾矿制备钛铁合金的方法

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AU3659695A (en) 1996-05-06
ZA958756B (en) 1996-05-15
GB2294255A (en) 1996-04-24

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