WO1996012047A1

WO1996012047A1 - Titanium and vanadium recovery process

Info

Publication number: WO1996012047A1
Application number: PCT/GB1995/002454
Authority: WO
Inventors: Peter Tyson; James Hamilton
Original assignee: Magmint Limited; Civelli, Carlo, Giuseppe
Priority date: 1994-10-17
Filing date: 1995-10-17
Publication date: 1996-04-25
Also published as: AU3659695A; GB9420937D0; GB2294255A; ZA958756B

Abstract

A process for the recovery or separation of titanium and vanadium from a slag comprising mixing the slag with an alkali metal salt, and roasting the mixture in oxygen.

Description

TITANIUM AND VANADIUM RECOVERY PROCESS

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 (Ti0₂) 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. Ti0₂ 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 Ti0₂ 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.

The largest application of 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.

Extraction processes are known for the production of vanadium or titanium from ores rich in these elements:

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.

Titaniferous magnetite ore containing approximately 13.1% Tiθ₂ 79.1% Fe₃0₄ and 1.6% V₂0₅ yields iron and V₂o₅ via a silica flux electric smelter process. The Ti0₂ together with the spent flux, is slagged off from the process for disposal as waste. No value is generated from the Ti0₂ slag, i.e. the Ti0₂ is discarded as waste.

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 Ti0₂ in the slag from the iron. However, 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.

In general those approaching metal extraction from vanadiferous titano-magnetite are faced with the following dilemma:

During recovery, where the vanadium reports largely to the slag in order to obtain an acceptable iron recovery, the levels of vanadium are too high for the production of Tiθ₂ slag. The upper impurity level of vanadium in the Ti0₂ extract is considered to be 0.6%. On the other hand where the operation is carried out under more highly reducing conditions there will be an increased and unacceptable vanadium concentration in the iron fraction together with the risk of an excessively viscous TiO-, slag with unacceptable levels of Ti0₃, TiC and unattackable Ti0₂.

There is therefore a desideratum to be able to follow through one of these processes to result in the recovery of both TiO.. and vanadium from titaniferous ore. There is a desideratum for an ore extraction process which resulted in the recovery of iron, Ti0₂ and vanadium from vanadiferous titano-magnetite. It is also a desideration to be able to upgrade Ti0₂ slag to remove conta ining quantities of vanadium. Accordingly, 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.

Thus, 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. 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. At this stage, 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. Preferably the particle size should not exceed 200μm. More preferably the particle size should be less than 45μm (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.

Preferably 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.

It is preferred that 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. Preferably 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. Preferably the oxygen should be supplied constantly throughout the process so as to maintain an oxygen rich atmosphere at least during loading and roasting.

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.

The leach liquid may be withdrawn from the mixture at this stage or may be recycled to the next charge. Preferably the leach liquid is withdrawn when a pre-determined vanadium concentration is reached. Following its removal from the slag 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|l is obtained. The vanadium may then be brought into solid form by the addition of ammonium chloride or ammonium sulphate to form solid ammonium metavanadate.

Preferably 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.

Preferably 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.

Thus 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 V₂0₅ content (<0.6%) , unattackable Ti0₂ (0.2%) and low Ti₂o. values (<0.1%) , all necessary for successful TiO. extraction.

This can be accompanied by 83% vanadium recovery. Transportation of the slag as pellets prevents dust handling problems. Having already been water/acid leached they are relatively impervious to any further accidental natural leaching (rain, etc) . The present process has been carried out on a starter slag the analysis of which is detailed below: Total Ti as TiO, 71.01%

Ti as Ti₂0₃ 12.0% Total Fe as FeO 21.61% Metallic Fe 1.2% Al as A1₂0 8.31% Si as Si0₂ 5.22% Ca as CaO 0.52% Mg as MgO 2.25% Mn as MnO 0.54% V as V₂0₅ 1.25%

The 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.

10 20 3C -0 ..NajCO^ Addi t i on In addition a slag of the above composition was subjected to the following alkali metal salts which resulted in a vanadium recovery of around 60%: Na₂CO₃/Na₂S0₄ Na₂C0₃/NaN0₃ and Na₂C0₃/K₂C0-,. Without oxygenation in a furnace the vanadium was not being fully oxidized thus resulting in a loss of V₂0₅.

Slags which had been crushed to different sized particles were investigated in an attempt to find a dust free product. In all cases after water leaching of 200(+75) , 70(+45) , 45 micron diameter particles, a fine powder was observed.

It was found that produced 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-,0₃ of 4%, Si0₂ 6% and Ti0₂ <1% in all water leach liquors.

Satisfactory water leaching time was found to be 30 minutes. Obviously on the plant the water leach liquor will be recycled until it is loaded to capacity with vanadium.

To further upgrade the slag a 20% sulphuric acid leach was investigated as follows:

A.20% acid leach on the original slag. B.20% acid leach on the roasted slag. C.20% acid leach on the water leached slag. The % recovery from the slag was as follows:

% RECOVERY V₂0₅ 3.0 85.0 3.0

A1₂0₃ 27.0 74.0 70.0

Si0₂ 12.0 63.0 57?0 Ti 0₂ 2 . 5 2 . 5 2 . 5

FeO 23.0 4.5 4.5

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) .

Predictably all of the vanadium was recovered using the acid leach on roasted material, the accompanying Si0₂, A1₂0₃ and Ti0₂ would make vanadium recovery impossible in the leachate.

Acid leaching of the water leached roast only served in removing SiO-.

Suitable furnace lining can also reduce the MgO.

The processes according to the invention described hereinbefore can be further improved by modification of individual steps including the following: -Mineral Beneficiation step:

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 Al₂0₃ is critical .

Pre-reduction of material prior to comminution and detailed magnetic separation results in a improved process. 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.

- Pre-reduction step:

The previously known pre-reduction stage was carried out on finely-ground magnetite. Both coal and liquid propane gase LPG are used as reductants. It was found that was most efficient using the fluidised bed and tube furnace. Coal was less efficient but is cheaper and easier to manage. Also known is to use only powdered material, which would have to be processed per fluidised bed or pelletised for use in a rotary furnace, with the consequent disadvantage of introducing more of the process-sensitive Al₂0₃ through the use of bentonite binder. Molasses is a more appropriate medium and can supply in-situ reductant.

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. 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% Ti0₂ magnetite. Similar efficiencies on 21% Ti0₂ feedstock are about 60% pre- reduced.

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 Ti0₂ values of about 27-29% with SiO- of <1% and A1₂0₃ in the 1.5% range. - Furnace Smelting step:

This concerns the process for the extraction of iron, vanadium and titanium from vanaderiferous titano-magnetite.

Excess reductant results in vanadium reporting to the metal phase and slags of >85% being produced. Some reduced oxides of titanium were found to be present in the slag. A deficiency of reductant resulted in vanadium reporting to the slag phase and a "vanadium- free" metal being produced.

Under the less reducing conditions, FeO was higher in the slag phase, with Ti0₂ 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).

In the case of the high-Ti0₂ slag, Ti0₂ content can be increased by this process from 85 to 92%, whilst Al₂0_j, Si0₂ and V₂0₅ would show corresponding decreases.

With the 66-75% Ti0₂ slag, processes described in Claim 1 would also be used to oxidise the slag and, in this case, to recover V₂0₅ from the slag, as well as removing FeO, A1₂0₃ and Si0₂ to acceptable levels.

In more detail 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 Al₂0₃, MgO and CaO, with 85% Ti0₂ or better, mostly in the form of lower Ti-oxides. 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.

With the axiom that metal produced under the relatively oxidising conditions is of sufficiently good quality to sell direct, these conditions were previously chosen. One advantage is that the liquidus temperature of slag and hence the system is lower because of the large amount of FeO reporting to the slag, which leads to greater stability of the furnace lining. A corollary is the loss of Fe to the slag phase which is ultimately discarded.

The consequence of adoption of these furnace conditions is that the slag is poor in Ti0₂ (about 66- 75%) , and V reports almost exclusively to the slag.

In contrast, the system can be run at conditions of over-reduction. This, however, does have some drawbacks. Of these, 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 Ti0₂ 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 However, if the furnace conditions are over-reducing with pre-reduced magnetite and char as feedstock, then point 2 above is negated. With a furnace freeze- lining, then 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. If required, 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. In accordance with this process 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 ₂0₅. Any vanadium present in the slag is in minimum quantities ( the majority having been extracted via the shaking ladle step) . In this situation the roasting and alkali metal salt process results in an upgraded Ti0₂ slag (approximately 92-94% Ti0₂) but there is no V recovery of any signficance. However, it is precisely the same process, as hereinbefore described with reference to the recovery of titanium and vanadium from slag (roasting with alkali metal salt) which enables the upgrading of the JTi slag to a composition of approximately 92-94% Ti0₂. Previously, it has not been possible to separate the vanadium from such a Ti rich slag. - Downstream slag upgrading step:

The slag produced from the relatively oxidising conditions of the route contains about 66-75% Ti0₂, some as lower oxides. The rest of the slag is made up of FeO (approximately 8-15%) with abundant A1₂0₃ (approximately 5%) , Si0₂ CaO, MgO and, of course, V₂0₅. The process as described hereinbefore according to the invention upgrades the slag into a more marketable form.

During the previously described (cold) leaching of the slag, with H₂S0₄, relatively little of the remaining, and unwanted A1₂0₃ is removed. However, leaching with H₂S0₄ at a temperature in the range of from 40 to 90°C removes up to 70% of the remaining Al₂0_j and Si0₂. The preferred temperature of the hot leach step is 60 to 85°C, more preferably around 80°C.

If the slag is heated in an oxidising furnace, then two advantages are identified. Firstly, the low Ti oxides are mostly converted to Ti0₂, which is highly soluble in the sulphate route. Hot leaching on this material then removes the V₂0₅ 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 H₂SO_A to cool pellets, by adding cool H₂S0₄ to the hot pellets, or even by heating a combination of cool pellets with cool H₂SO .

In a process where the majority of the V has been removed by the shaking ladle step, then a water leach step (previously used to leach the V) is not required. The high-temperature (up to and around 85%) Ti0₂ 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 A1₂0₃ 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% Ti0₂ is possible by optimization of leach conditions. - Final products

After metal treatment (soft-blow to remove Vanadium and TiC, desulphurization and addition of Ferrosilicon) , then 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 εynrutile in composition, with approximately 92-94% Ti0₂ as higher oxides, 2-3% A1₂0₃, <1% Si0₂ 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 V₂0₅, and that oxide is also available from the final clean-up of the slag. It has been proposed that V₂0₅, Fe and Al can be combined by exothermic reaction into FeV.

Embodiments of the present invention will now be described by way of example only with reference to the accompanying drawings of which:

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; and

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.

Referring to Fig. 1: 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) is balled milled to particles of approximately 45 microns and these are stored in a slag bunker. 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.

Referring to 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.

Referring to Fig. 3: 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) . By the term benefication 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.

Referring to Fig.4: 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 75μm. 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% Ti0₂ 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 Ti0₂ 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° Ti0₂) 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 V₂0₅ (which may be precipitated) , and removing the majority of Si0₂, A1₂0₃, CaO and MgO impurities. The final slag composition comprises approximately 92 to 94% Ti0₂.

It will, of course, be appreciated that 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.

Claims

CLAIMS :

1. 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 preferably at a temperature in the range of from 900 to 1200°C and adding a leaching liquid to the mixture from which the vanadium is recovered.

2. A process as claimed in claim 1 further comprising crushing and optionally milling the slag into a reduced size slag before mixing with the alkali metal salt.

3. A process as claimed in claim 2 wherein after the slag is crushed oversized lumps containing iron prill are removed and discarded.

4. A process as claimed in any one of claims 1 to 3 wherein the slag and the alkali metal salt are mixed in the ratio of 5:1 weight by weight.

5. A process as claimed in any one of claims 1 to 4 wherein before roasting the mixed slag and the alkali metal salt are wetted, pelletized and dried.

6. A process as claimed in any one of claims 1 to 5 wherein the roasting is carried out at approximately 1125°C.

7. A process as claimed in any one claims 1 to 6 wherein the roasting is continued for approximately 30 minutes.

8. A process as claimed in any one of claims 1 to 7 wherein during roasting the oxygen is provided as an upward flow in a vertical roaster.

9. A process as claimed in any one of claims 1 to 8 wherein after roasting the mixture is quenched at approximately 90°C in the leach liquid.

10. A process as claimed in any one of claims 1 to 9 wherein the leach liquid is recycled to the next charge until a predetermined vanadium concentration is reached.

11. A process as claimed in any one of claims 1 to

10 wherein the leach liquid is treated with aluminium sulphate to remove silica and alumina.

12. A process as claimed in any one of claims 1 to

11 wherein the leach liquid is water.

13. A process as claimed in any one of claims 1 to

12 wherein after removal from the mixture the leach liquid is evaporated to a vanadium concentration of approximately 70g|l.

14. A process as claimed in any one of claims 1 to

13 wherein the pH of the leach liquid is adjusted by approximately 8 by the addition of alkali and ammonium metavanadate is precipitated by the addition of ammonium chloride, ammonium sulphate or a mixture thereof.

15. A process as claimed in any one of claims 1 to

14 wherein the remaining slag mixture is further leached in dilute sulphuric acid, washed and dried to produce upgraded titanium dioxide slag.

16. A process as claimed in claim 15 wherein the leaching is carried out in the temperative range of 40 to 95°C

17. A process as claimed in any one of claims 1 to

15 wherein the alkali metal salt is sodium carbonate, sodium hydroxide, sodium sulphate or a mixture of two or more thereof.

18. A process as claimed in any one of claims 1 to

16 wherein the alkali metal salt is sodium carbonate and sodium sulphate in a ratio of 1:1 by weight.

19. A process for the recovery of iron, vanadium and titanium from vanadiferous titano-magnetitie 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 the vanadium by a process as claimed in any one of claims 1 to 16.

20. A process as claimed in claim 17 wherein before the first pre-reduction step the magnetite is crushed and/or ground to a reduced size magnetite.

21. A process as claimed in claim 19 or claim 20 wherein the smelting process is carried out in an electric arc furnace.

22. A process as claimed in claim 19 or claim 20 wherein the smelting process is carried out in a furnace under reducing conditions.

23. A process for upgrading a titanium slag to reduce vanadium impurities comprising mixing the slag with a alkali metal salt and roasting the mixture in oxygen preferably at a temperature in the range of from 900 to 1200°C

24. A process as claimed in claims 22 in combination with 1 or more of the features described in claims 2 to 8 or 15 to 18.

25. Vanadium and/or titanium or salts thereof when recovered by a process as claimed in any one of claims 1 to 24.

26. Vanadium, titanium or salts thereof or iron when recovered by a process as claimed in any one of claims 19 to 22.