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TITLE
METHODS FOR SEPARATION OF TITANIUM FROM ORE
Field of the Invention
The present invention relates to methods for the separation of titanium
from ore, especially iron-bearing ore e.g. ilmenite ore. In embodiments of the
invention, the method relates to the recovery of titanium tetrahalides,
especially titanium tetrachloride from solutions. In further embodiments, the
invention relates to recovery of titanium metal from such ore.
Background of the Invention
Many processes are known for the recovery of titanium dioxide from
ores. Ilmenite, which contains mainly titanium oxide and iron oxide values,
often is employed in such processes. The majority of processes for the
recovery of titanium dioxide from ores involve digestion of the ore in a
mineral
acid, such as hydrochloric acid or sulphuric acid, to remove at least the
titanium values from the ore. In such processes, however, the purity of the
titanium dioxide produced is about 90-95%, and hence further purification
procedures may be required to produce a pigment grade product, which adds
considerably to the cost. Many of the further purification procedures involve
techniques that are environmentally unacceptable without extensive
procedures to treat various solutions and solids obtained. Such treatment
processes tend to be costly.
Processes for the recovery of titanium dioxide from ilmenite in high
purity and high yield are known. One such process is described in U.S.
Patent No. 3,903,239 of S.A. Berkovich, which discloses a process which
comprises contacting ilmenite, or a concentrate thereof, in particulate form
with concentrated hydrochloric acid at a temperature of about 15-30°C
to
solubilize and leach from the ore at least 80%, preferably at least 95%, of
the
iron and titanium values. The leaching operation may be carried out over an
extended period of time, typically from 3-25 days, depending on the technique
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employed and the quantity of iron values to be recovered. Leaching
techniques include counter-current flow or the use of closed cycle loops in
which hydrochloric acid is continuously passed through a bed of the ore. The
leaching operation is exothermic, and the reactants are maintained in a
temperature range of 15-30°C by cooling, if necessary.
The ilmenite or similar ore used in the process may be treated as such
or may be beneficiated to form a concentrate in any desired manner. Ilmenite
generally contains Ti02.Fe0 with varying amounts of Fe203 and gangue
materials, usually silicates, alumina, lime and magnesium. Beneficiation may
be employed when the ore is of low Ti02 content.
The ore or concentrate may be pre-treated prior to contact with the
concentrated hydrochloric acid to increase the rate of dissolution of the
titanium and iron values during the leaching step. Such pre-treatment may
include an initial oxidation at elevated temperature, such as from 600-
1000°C,
in the presence of air and/or oxygen to split the Ti02.Fe0 followed by
reduction of at least part of the iron oxide with carbon or carbon monoxide.
This is a smelting step, with slag from the smelting step being fed to the
leaching step and pig iron being marketed.
Subsequent to the leaching step, it is necessary to convert any ferric
iron in the solution to ferrous iron, which is typically achieved by reduction
of
the ferric iron in the leach liquor with a gaseous reducing agent e.g. sulphur
dioxide. The conversion of ferric iron to ferrous iron in this manner is
essential in view of the affinity of titanium dioxide for ferric iron and the
difficulty in separating ferric iron from titanium dioxide.
The solution of titanium chlorides and ferrous chloride which is thus
obtained, and which may contain minor quantities of gangue metal chlorides,
typically calcium and magnesium materials, is then mixed with water to cause
hydrolysis of the titanium chlorides. A seeding amount, generally about 1-2%,
by weight of the titanium oxyhydrate to be precipitated (Ti02.3H20) is
included
in the mixture. Titanium oxyhydrate precipitates from the mixture. The
hydrolysis is carried out using a quantity of water at least sufficient to
precipitate substantially all of the titanium values from the solution but
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insufficient to cause precipitation of other metal oxides or hydroxides. The
titanium oxyhydrate that is precipitated from the mother liquor is then washed
substantially free of entrained mother liquor and dried. The washed
precipitate
is converted at elevated temperature, typically 700-1000°C, in the
presence of
air and/or oxygen into the anatase or rutile form of titanium dioxide.
Alternative methods that require less treatment of solution and solids to
ensure environmental acceptance and/or are less expensive in achieving
environmental acceptance, as well as producing titanium and other products
of high value e.g. high purity, are required.
Summaryr of the Invention
A method for the separation of titanium, and for production of titanium
metal, from titanium-bearing ore that involves a reduced number of steps has
now been found.
Accordingly, one aspect of the present invention provides a method for
the separation of iron values from titanium-bearing ore, comprising the steps
of:
a) leaching said ore, or a concentrate thereof, with an aqueous
solution of a hydrogen halide at a temperature of at least 90°C;
b) separating solids from the leach solution obtained in step (a), to
provide a leachate solution;
c) subjecting the leachate solution to extraction with an immiscible
organic phase that selectively extracts iron values into said organic phase,
titanium values in the leachate solution selectively remaining in the aqueous
leachate solution, said immiscible organic phase having an extractant
selected from the group consisting of phosphoric, phosphonic and phosphinic
acids, and oxides thereof.
Another aspect of the invention provides a method for the production of
titanium metal from titanium-bearing ore, comprising the steps of:
a) leaching said ore or a concentrate thereof with an aqueous
solution of a hydrogen halide at a temperature of at least 90°C;
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b) separating solids from the leach solution obtained in (a), to
provide a leachate solution;
c) subjecting the leachate solution to extraction with an immiscible
organic phase having a boiling point that differs from the boiling point of
titanium halide formed in step (a) in the leachate by an amount that permits
separation of titanium halide from the organic phase by fractional
distillation,
said organic phase containing the titanium halide and being stable with
respect to the titanium halide; and
d) stripping titanium halide from the organic phase obtained in step
(c) by heating to volatilize the titanium halide and effect separation from
the
organic phase.
A further aspect of the invention provides a method of separating a
titanium halide from a concentrated aqueous solution of the titanium halide,
said titanium halide being in a concentration such that the titanium halide is
substantially stable in said aqueous solution, comprising:
a) admixing said aqueous solution with an organic phase having a
boiling point that differs from the boiling point of the titanium halide by an
amount that permits separation of titanium halide from the organic phase by
fractional distillation;
b) separating the organic phase so obtained from the aqueous
solution, said organic phase containing titanium halide; and
c) heating the organic phase and stripping titanium halide
therefrom.
Yet another aspect of the invention provides a method for the
separation of titanium from a titanium-bearing ore, said ore containing iron,
comprising the steps of:
a) leaching said ore, or a concentrate thereof, with an aqueous
solution of a hydrogen halide in the presence of an oxidising agent; and
b) effecting a separation of titanium dioxide obtained in step (a)
from said solution.
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A further aspect of the invention provides a method of forming a titania-
rich slag from a titanium-bearing ore that contains iron, comprising the steps
of:
a) calcining the ore under oxidizing conditions to eliminate sulphur
from said ore, said calcining being carried out at a temperature of at least
1200°C;
b) subjecting the hot calcined ore of step (a) to reducing conditions
in the presence of CO;
c) transferring the hot reduced calcined ore obtained in step (b) to
a smelting step;
controlling the reducing conditions and the smelting step to obtain pig
iron and a titanic-rich slag with a predetermined iron content.
Another aspect of the invention provides a method of forming a titania-
rich slag from a titanium-bearing ore that contains iron, comprising the steps
of:
a) calcining the ore under oxidizing conditions to eliminate sulphur
from said ore, said calcining being carried out at a temperature of at least
1200°C;
b) subjecting the hot calcined ore of step (a) to reducing conditions
in the presence of CO;
c) transferring the reduced calcined ore obtained in step (b) to a
leaching step in an aqueous solution of weak sulphuric acid;
controlling the reducing and leaching conditions to obtain a titanic-rich
material having less than 5% by weight of the iron content of the ore.
A further aspect of the invention provides a method for the separation
of iron and titanium values from an iron/titanium ore, comprising the steps
of:
a) leaching said ore, or a concentrate thereof, with an aqueous
solution of a hydrogen halide in the presence of an oxidising agent; and
effecting a separation of titanium dioxide obtained from said solution.
b) subjecting the aqueous leach solution so obtained to extraction
with an immiscible organic phase that selectively leaches iron values into
said
organic phase.
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Yet another aspect of invention provides a method for the separation of
iron values from titanium-bearing ore, comprising the steps of:
a) leaching said ore, or a concentrate thereof, with an aqueous
solution of a hydrogen halide;
b) separating solids from the leach solution obtained in (a), to
provide a leachate solution;
c) subjecting the leachate solution to extraction with an immiscible
organic phase that selectively extracts iron values into said organic phase,
titanium values in the leachate solution selectively remaining in the aqueous
leachate solution; and
d) subjecting the aqueous leach solution obtained in step (c) to
steps to separate Ti02 therefrom.
Detailed Description of the Invention
Processes for the recovery of titanium dioxide from ilmenite, with high
purity in high yield, are known. Techniques for treating the ilmenite ore,
optionally to form concentrate and/or for beneficiation of the ore are known.
In some instances, it is possible to treat the ore or concentrate with
concentrated hydrochloric acid solution to effect a leaching of titanium
values
from the ore or concentrate. In other instances, it is necessary or desirable
to
subject the ore or concentrate to a smelting step in the presence of carbon
and/or fluxing agents, and to then separate a slag from the smelting process
which is then subjected to the leaching step.
One aspect of the present invention is directed to the step of recovery
of titanium values from the leaching solution. The leaching solution is a
mixture of aqueous hydrochloric acid containing titanium values, and other
soluble material and solid materials, particularly residues of the concentrate
and/or slag from which the titanium values have been leached. A liquid/solid
separation step is conducted, to separate a leachate solution from solids.
The leachate so obtained is treated, according to an aspect of the
present invention, with an organic phase. The titanium values in the leachate
are in the form of titanium halide, especially titanium tetrahalide which, if
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hydrochloric acid is used in the leaching step, will be titanium
tetrachloride.
In one aspect of the invention, the organic phase is selected so that
iron values are selectively separated into the organic phase. Thus, an
organic/aqueous separation is effected, with the iron values being in the
organic phase and titanium values remaining in the aqueous phase.
Preferably, iron values are separated almost to the exclusion of other values
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in the leachate solution, or with values readily separated therefrom, so that
iron products especially iron oxides, may be obtained in high purity.
Vanadium and other metal values that may be present in the leachate are
preferably retained in the aqueous phase.
Examples of the organic phase are phosphoric, phosphoric acid and
phosphinic acids, esters oxides thereof. Specific examples are tri-n-butyl
phosphate and di-2-ethylhexyl phosphoric acid. The organic phase preferably
contains a diluent e.g. a hydrocarbon, an example of which is a kerosene.
The organic phase may be stripped from the iron values and recycled.
The iron values may then be subjected to pyrohydrolysis, or other steps, to
recover iron e.g. as iron oxides. Preferably HCI is obtained as a by-product
which is recycled to the step of leaching of ore or concentrate described
above.
In a second aspect, the organic phase is selected such that the
titanium halide is soluble in the organic phase. Moreover, the organic phase
is selected such that the organic phase and titanium halide may be separated
by fractional distillation. The organic phase may be selected to have a higher
or lower boiling point that the titanium halide. The preferred titanium halide
is
titanium tetrachloride. In embodiments of the invention, the boiling point of
the organic phase differs by least 50°C, and especially at least
75°C, from the
boiling point of the titanium halide. The organic phase must be immiscible
with the aqueous solution, such that it forms a second layer so that
separation
may be effected.
The titanium halide is extracted from the aqueous solution into the
organic phase, to effect removal of the titanium halide from the aqueous
solution. Such extraction may be carried out in a continuous operation or in a
batch operation.
The organic phase may be, for example, a crown ether, phosphine
acid, ester or oxide, or tertiary or quarternary ammonium salt.
The organic phase containing the titanium halide is separated from the
aqueous solution and from any solid matter, and is then subjected to a step to
separate the titanium tetrachloride. In the separation step, the organic phase
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containing the titanium halide is heated to effect the separation of the
titanium
halide. This is preferably accomplished by volatilization of the titanium
halide,
especially if the halide is chloride although for any particular tetrahalide
the
organic phase may be selected to effect volatilization of either the titanium
tetrahalide or the organic phase. In addition, the organic phase should be
selected so that it has a flash point that is acceptable under the operating
conditions, preferably a flash point above the temperature used in separation.
The organic phase needs to be stable with respect to the aqueous solution
and to the titanium tetrahalide at the operating conditions.
The aqueous solution remaining after extraction may be subjected to
known procedures for recovery of iron or other metal values or procedures
described above, and for recovery and recycle of acid used in the leaching
step. For example, iron oxide (Fe203) may be recovered and the acid e.g.
HCI, recycled. The organic phase is preferably recycled back to the extraction
step, and reused.
The titanium tetrahalide may be subjected to purification steps, if
necessary. However, if the titanium tetrahalide is volatilized, it may be of
acceptable purity for many end-uses. The titanium halide may be used as
such or subjected to further processing steps e.g. to form titanium metal.
Techniques for the conversion of titanium tetrahalide and titanium dioxide to
titanium metal are known.
In embodiments of the invention, essentially all of the iron of the feed
material i.e. titanium-bearing ore or concentrate is dissolved by the HCI.
Thus, sufficient HCI has to be provided. In order to minimise the amount of
HCI required in the process, the iron chloride (e.g. H+FeCI-4 or a chloride of
iron e.g. FeCl3) produced in the process is subsequently subjected to
pyrohydrolysis to regenerate the HCI. The iron is converted into an iron oxide
product (Fe203). By controlling the composition of the iron chloride solution
before subjecting it to pyrohydrolysis, it is possible to produce a high grade
iron oxide suitable for use in pigment production. If this is not economic,
the
disposal of iron oxide becomes an environmental problem. In such instances,
it is advantageous to remove the iron and upgrade the titanium ore before
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subjecting it to treatment with HCI. This alternative also has the advantage
of
decreasing the size of the hydrometallurgical plant for a given production of
titanium.
In another aspect of the present invention, the ore or concentrate is
leached in aqueous solution in the presence of an acid and an oxidizing
agent. A variety of oxidizing agents may be used, including air, hydrogen or
other peroxides, or sodium or other perchlorates. The oxidizing agent should
be selected to minimize any contamination of the solution with cations that
have an adverse effect on other process steps.
In the aqueous solution, the titanium is converted to titanium dioxide
and the iron is solubilized. The acid is preferably a hydrogen halide,
especially HCI. If the acid is HCI, the concentration of acid may be
controlled,
in the presence of the oxidizing agent, to convert iron into H+FeCI-4, which
is
soluble in the aqueous solution. Subsequently, liquid/solid separations may
be effected, to separate Ti02 including separation of Ti02 from tails from the
aqueous solution. The aqueous solution may be treated for recovery of HCI
and iron e.g. as Fe203.
One of the concentration alternatives for the treatment of the ilmenite
ore or physically beneficiated ilmenite is the production of a titanium-rich
slag
and pig iron from it, with the titanium-rich slag being subjected to HCI
leaching
for titanium recovery and the iron being sold as a foundry-grade pig iron.
The processing of ilmenite to produce titanic slag and pig iron is
known. In such processing, the concentrate is calcined in a rotary kiln under
oxidising conditions at about 1200-1300°C to eliminate sulphur in the
ilmenite
concentrate. The product is cooled and fed to an electric furnace with
coal/coke as reductant to reduce the iron and produce a molten slag and pig
iron. The electric furnace smelting takes place at about 1650°C.
Disadvantages of the current processing method include (i) the energy in the
calcine from the rotary kiln is lost and the product must be re-heated to the
smelting temperature in the electric furnace with electric power; (ii) the
reduction in the electric furnace produces a lot of CO gas, which often
results
in the foaming of the slag and process control is difficult; and (iii) the
reduction
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of the iron oxide by carbon to produce CO is endothermic and the heat
required is supplied by electricity.
The production of electricity from fuel is typically energy inefficient, to
the extent of about 30% conversion, and therefore the process is very energy
intensive.
In aspects of the present invention, the use of energy is reduced. In a
preferred embodiment, the reduction of the calcine is carried out prior to the
electric furnace smelting. This may be carried out in a second reducing
section of the calcining kiln or a separate kiln following the calcining kiln.
The
reduction is done by reducing gases produced by the partial reduction of CO
(produced in the electric furnace) and/or by partial combustion of fuel. The
reduced calcine so produced is transferred hot from the reduction kiln to the
electric furnace, with any additional reductant if required. This saves the
energy lost in cooling the calcine and improves the efficiency of use of fuel
for
the iron oxide reduction. The smelting of reduced calcine requires less
electric energy, produces very little gas in the electric furnace and makes
the
electric furnace easier to control. The reduction in the reduction step and
the
electric furnace are controlled to provide a desired iron level in the slag.
This
controls the slag melting temperature. Depending on the composition of the
ilmenite concentrate, the smelting may be carried out at lower temperatures
e.g. 1550°C, thereby conserving more energy compared to the current
processing.
Alternatively, the process of smelting and production of molten titania-
rich slag and pig iron in the final step of the above process may be replaced
by a leaching process for removing the reduced iron produced in the reduction
step. This may be carried out by dissolving it in a weakly acidic chloride
solution, with aeration, to dissolve the iron and leave a titania-rich oxide
product suitable for HCI leaching, as described herein. The advantage of this
approach is that the reduction can be carried out at lower temperature e.g.
about 800°C instead of about 1000°C and the energy of smelting
in the
electric furnace is conserved. In addition, in this embodiment, it is possible
to
eliminate over 95% of the iron and minimise the iron fed to the HCI leaching
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process. In comparison, there is about 90% iron removal in the smelting
route, as some iron has to be left behind to provide a fluid slag in the
smelting
step. This minimises the usage of HCI and the subsequent regeneration of
the HCI for re-use in the process.
In the production of titanium and Ti02 pigment, it is necessary to
produce titanium chloride by chlorination of titanium dioxide-containing
materials e.g. titania slag and rutile concentrate, at about 900°C in
the
presence of coke. By using the present process, the titanium chloride may
be produced by leaching the titania containing material with HCI followed by
extraction and separation of titanium chloride. The high temperature
chlorination is replaced by lower temperature operations and avoids the
formation of environmentally unacceptable dioxins which can form in the high
temperature chlorination. In addition, the purity of the titanium chloride
produced in the present process will be improved and will require less
purification or even no purification.
In various aspects of the invention, there is provided a process in which
titanium-bearing ore or concentrate, generally after having been subjected to
a smelting step, is subjected to a leaching step using aqueous hydrochloric
acid. A leach solution containing titanium and iron values, and other metallic
values depending on the particular ore, is obtained. A liquid/solid separation
step is conducted. The solids may be subjected to other separation steps but
generally will be gangue.
In preferred aspects of the invention, the leachate solution is subjected
to extraction with organic phase to extract iron values, as described above.
Such phase may contain 100-200 g/1 of iron, or more. Examples of the
organic phase are tri-n-butyl phosphate and di-2-ethylhexyl phosphoric acid,
with other examples being described above. The organic phase with iron
values is then subjected to steps to separate and recover the organic phase,
which is recycled to the step of extraction of the leachate solution. Iron
values, which are in the form of chlorides are recovered, especially by
pyrohydrolysis to yield iron oxides and HCI. The HCI is recycled to the
leaching of the ore or concentrate. Iron values of high purity may be
obtained.
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The leached aqueous solution or raffinate may be treated to reduce the
concentration of metallic impurities prior to treatment for separation of
T;02.
For example, the raffinatE~ may be treated to separate vanadium and other
metallic values, depending on the particular ore, especially by precipitation,
to
provide a raffinate rich in titanium, and preferably with high-purity titanium
values. The titanium values are in the form of the chlorides viz TiCl4, which
may be subjected to steps to form Ti02, which is recovered. High purity Ti02
may be obtained, which may be of sufficiently high purity for use as such.
As an alternative, t:he raffinate rich in titanium values may be subjected
to further extraction, using an organic phase that is immiscible in water and
which has a boiling point that differs from the boiling point of the titanium
value, e.g. titanium tetrachloride by an amount to permit fractional
distillation.
The titanium value is extracted from the aqueous solution of the raffinate
into
the organic phase and then recovered by distilling or flashing off either the
organic phase or the titanium value, depending on the respective boiling
points. Titanium metal may then be recovered from the titanium halide.
The by-products of such a process are minimized, and may be treated
by known but relatively simple techniques.
An alternative separation of titanium values is to separate the titanium
as Ti02 directly from the leaching of the ore or concentrate, by leaching in
the
presence of an oxidizing <agent and separating the Ti02 formed from the
solution and from other solids therein.
The present invention provides methods for the separation of titanium
from titanium-bearing ores, especially ilmenite. In particular, the invention
provides methods for production of titanium tetrahalides, especially titanium
tetrachloride, and Ti02 with improved purity and/or such that related steps in
an overall process, including recovery and recycle of materials, may be
simplified and be more cast effective. In particular, the volumes of liquid
and
solids that must be handled in the overall separation process may be reduced,
and associated hardware may be reduced in size. Such improvements may
be of significant economic; benefit.
The present invention is illustrated by the following examples.
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EXAMPLE I
A sample of a concentrate of a titanium-bearing ore of a particle size
such that 59% by weight would pass through a 100 mesh screen, was
subjected to HCI in an amount estimated to be 100% of the stoichiometric
amount of chloride required for the amount of titanium in the concentrate. The
temperature was 95°C.
The amount of concentrated HCI was 541.5 g, for 100 g of concentrate.
After a period of 2 hours, 10 g of NaCl03 were added.
The total treatment time was 3 hours.
The calculated amount of titanium in the samples was 18.5 g, and the
amount of iron was calculated to be 39.8 g.
The resultant solution was filtered, washed with an acid solution (HCI)
and then with water. The solution were analyzed. The results obtained were
as follows:
CA 02289967 1999-11-17
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O I~
M O M M r-
N
N N ~ T ~ ~ O
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U O M
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_ ~
H O O O M O O)
~7
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O O
~ 00 t!7
p LL M M N O O O M
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'
~ j ~ O O CD
~ _ ~ ~ 00
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R
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'- M ~- N O d'
O O
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CO M f~
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w C
O
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cn ~ +, ~ - ca
H LL LL
CA 02289967 1999-11-17
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This example shows the use of oxidant in the recovery of titanium
metal values, as Ti02, from a titanium bearing ore. Titanium is effectively
separated from iron; as indicated by the high iron content of the filtrate and
the low iron content of the fine tails.
EXAMPLE II
Initial test work on extraction of synthetic titanium chloride solution
showed that trialkyl phosphine oxide (CyanexT"" 923) could be a good
candidate for titanium extraction. Tests were then conducted on leach solution
produced by leaching of ilmenite concentrate. These tests showed that iron
preferentially loads into trialkyl phosphine oxide from a solution containing
titanium and iron. In a typical test, pregnant leach liquor diluted with an
equal
volume of distilled water, assaying 30.1 g/1 Titanium and 30.0 g/1 Iron, was
contacted with HCI pre-conditioned 100% trialkyl phosphine oxide. In this
test,
sodium chlorate (NaCl03), was added to the aqueous phase to raise the EMF
of the pregnant leach liquor from 406.3 my to 567.3 my to aid the extraction.
Employing an organic to aqueous ratio of 1:2 and a 30 minute contact time at
25°C for extraction yielded an aqueous raffinate assaying 27.5 g/1
Titanium
and 2.15 g/1 iron. This represented an extraction efficiency of 93% for iron
and
8.6% for titanium.
This example illustrated that iron could be extracted first from the leach
solution.
EXAMPLE III
Test work on different extractants showed that tri-n-butyl phosphate
(TBP) is a good candidate for iron extraction from the leach solution prior to
titanium extraction. Large volumes of titanium-containing aqueous leach
solution prepared for extraction testing were repeatedly contacted with 100
vol. % TBP at 40°C at a 2:1 or 3:1 organic/aqueous phase ratio. After 6
to 8
full contacts, the resulting aqueous raffinate assayed <50 ppm iron. This
confirmed the feasibility of extracting iron from leach solutions. The
titanium is
not extracted by TBP and therefore it is possible to selectively remove iron
from leach solution. In these tests it was seen that when using 100% TBP the
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iron containing organic is viscous and it was therefore considered preferable
to use TBP with a diluent.
A 20% TBP solution in CF-231 kerosene diluent was used to determine
the extraction and stripping isotherms. The organic phase was pre-
conditioned with 200 g/1 hydrochloric acid prior to extraction of the iron.
The
feed aqueous used for these tests analyzed as follows (ppm): Ti 33700, Fe
76000, AI 210, B <5, Ba <1, Ca 150, Cd <5, Co <5, Cr 180, Cu 36, K <100,
Mg 1450, Mn 160, Mo 5, Na 2400, Ni 4.8, Pb 30, and Zn 14.
Varying phase ratios of organic and aqueous phases were tested with
beaker/magnetic stiffer bar contacts at 40°C for an interval of 10
minutes. The
observations in these tests were as follows:
1. 20 volume % TBP in CF-231 diluent organic phase saturates at about
22 g/1 of Fe.
2. The organic-aqueous separation is fast and is complete in about 1
minute.
3. A contact time of 10 minutes provides for the transfer from the aqueous
to organic phases.
4. These results show that an aqueous pregnant leach solution assaying
75 g/1 iron, may be extracted using 20 volume percent Tri-Butyl
Phosphate in CF-231 diluent in two or three stages at 40°C
contacting
5 parts Aqueous and I part Organic for 10 minutes.
EXAMPLE IV
After iron extraction, the TBP is stripped to remove the iron and the
organic phase should be returned to the extraction step. Iron extraction was
theoretically shown to require five or six stages using 20 volume % TBP. A
stripping isotherm was charted after contacting varying phase ratios of
organic
and aqueous at 40°C for 5 minute contact intervals. Leach pregnant
solution
was used to fully load 20 volume percent TBP in 80 volume percent CF-231.
Repeated contacts of fresh aqueous with one organic volume produced a
loaded organic solution assaying about 22 g/1 iron. This organic solution was
stripped with a 13.7 g/1 hydrochloric acid solution.
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Varying phase ratios of organic and aqueous solution were tested with
beaker/magnetic stiffer bar contacts at 40°C for an interval of 5
minutes. The
observations and analytical results are tabled below:
1. Iron loaded organic assaying 22 g/1 iron, can be stripped efficiently with
mild HCI at 13.7 g/1 concentration in six stages at 40°C contacting 1
volume of aqueous with 4 volumes of loaded organic.
2. The organic tested viz. 20 volume percent Tri-n-Butyl Phosphate in CF-
231 diluent, can be stripped in a six stage counter-current contact to a
concentration of 106 ppm iron. This stripped organic can be recycled
for extracting iron.
3. By manipulating the concentration of TBP in CF-231 diluent and by
increasing the contact phase ratio to 10:1 organic/aqueous, pregnant
strip solutions can be achieved higher than the 75 g/1 Fe obtained in
this isothermal testwork. An example based upon theoretical loading of
60 volume % TBP would be an iron concentration of 250 g/L achieved
in the pregnant strip solution.
An example of one run is shown in the attached Table 2.
A bulk TBP raffinate was produced in a series of extraction contacts of
pregnant leach liquor with barren TBP organic. This raffinate analyzed as
follows (the composition of the feed aqueous is given in Example III):
AI 500 ppm, B <5 ppm, Ba <1 ppm, Ca 230 ppm, Cd <5 ppm,
Co <5 ppm, Cr 340 ppm, Cu 44 ppm, Fe <50 ppm, K < 100 ppm,
Mg 2520 ppm, Mn 390 ppm, Mo <5 ppm, Na 8030 ppm, Ni 16 ppm,
Pb 39 ppm, Ti 55600 ppm and Zn 15 ppm.
The stripped iron solution can be used to regenerate the HCI and also it could
be used to produce a value added iron oxide product.
CA 02289967 1999-11-17
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CA 02289967 1999-11-17
-19-
EXAMPLE V
The use of diethyl phosphoric acid (DEHPA) with tributyl phosphate
(TBP) to remove vanadium was tested. A larger volume (400m1) of feed TBP
raffinate was treated under sulphur dioxide gas for one hour at 60°C.
The pre-
reduced aqueous solution was then contacted with HCI pre-conditioned
TBP/DEHPA organic extractant for 1.5 hours at 60°C while
maintaining the
S02 bubbling. The organic/aqueous emulsion was then allowed to settle and
separated. The resulting aqueous raffinate had evaporated from the feed
volume of 400 ml to 286 ml.
A multi-element analysis was carried out to determine vanadium
transfer into the solvent. Feed aqueous and resulting TBP/DEHPA raffinate
assays were as follows:
Aqueous Feed TBP/DEHPA Raffinate
AI - 560 ppm AI - 1220 ppm
Cr - 370 ppm Cr - 610 ppm
Cu - 48 ppm Cu - 1.5 ppm
Fe - 82 ppm Fe - 11 ppm
Mg - 2810 ppm Mg - 4140 ppm
Mn - 460 ppm Mn - 730 ppm
Ni - 50 ppm Ni - 34 ppm
Ti - 31000 ppm Ti - 36500 ppm
V - 360 ppm V - 104 ppm
Zn - 16 ppm Zn - 1.1 ppm
Vanadium was extracted into the aqueous phase at an efficiency of
79.3% in a single extraction. Additionally, iron, nickel, copper and zinc were
loaded into the TBP/DEHPA solvent.
EXAMPLE VI
Precipitation of Ti02 was also investigated under gas-reducing
conditions in conjunction with distilled water to induce titanium hydrolysis.
Tri-
CA 02289967 1999-11-17
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n-butyl phosphate raffinate assaying 74.9 g/1 Ti and <5 ppm Fe was
agitated/heated to at least 60°C while bubbling in 0.5 I/min S02 gas
for a 2
hour interval. Once the temperature was increased to 95°C, while
maintaining
S02flow, three times the volume of TBP raffinate were added as distilled
water. The one-hour reaction period at 95°C yielded a white wet-cake
precipitate. Thorough re-pulp washing and oven drying resulted in a light
brown solid which assayed as follows:
Fe 1.28 wt
Cr 0.364 wt.%
Si 0.246 wt.%
Cu 0.094 wt.%
Mg 0.087 wt.%
N i 0.086 wt.
Na 0.052 wt.%
V 0.031 wt.
P 0.030 wt.
Zn 0.029 wt.
Mn 0.028 wt.%
Pb 0.023 wt.%
All other metals were reported as <0.01 wt.% concentration. The
precipitation barren solution analyzed only 2.9 ppm titanium, indicating a
precipitation efficiency of greater than 99%. Other metals detected in the
precipitation barren aqueous were follows:
AI 300 ppm
Ca 220 ppm
Co 4.0 ppm
Cr 31 ppm
Cu 11 ppm
Mg 940 ppm
CA 02289967 1999-11-17
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Mn 140 ppm
Na 70.9 g/L
Ni 5.2 ppm
V <1 ppm
Zn <2 ppm
Example VII
The production of a Titanium Dioxide product of higher purity, at about
99 or 99.9% Ti02, can be achieved by removing all metal contaminants from
the leach solution or by selective precipitation. Iron removal has already
been
demonstrated in Example V with the use of tri-n-butyl phosphate organic
phase.
Titanium 64.7 g/1
Iron <50 ppm
Aluminum 650 ppm
Chromium 260 ppm
Copper 46 ppm
Magnesium 1790 ppm
Manganese 310 ppm
Nickel 10 ppm
Vanadium 47 ppm
The more difficult metal contaminants to extract from aqueous
hydrochloric media are vanadium and chromium. The use of a reducing agent
with the TBP raffinate converts vanadium into the solvent-extractable V4+
oxidation state. Chromium may then be separated from titanium by
precipitating titanium dioxide at a low pH before chromium, aluminum,
magnesium and manganese are precipitated.
In this example, a small aliquot of TBP raffinate was heated to
70°C
under magnetic stirrer bar agitation. Sulphur dioxide gas was bubbled into the
aqueous solution for 30 minutes. To extract V4+, an equal volume of 50/50
CA 02289967 1999-11-17
-22-
TBP(100%)/DEHPA(100%) was added to the aqueous solution. Agitation
with S02 gas was continued for 1 hour at a temperature of 75°C. This
low pH
aqueous/organic emulsion qualitatively showed an organic appearance
change from a colourless organic to a light golden yellow organic.
Additionally, the heating for an extended period of time caused a pure white
titanium dioxide precipitate to separate out of the aqueous solution.
Although precipitation efficiency of titanium was not established in this
example, the resulting multi-element assay revealed a very pure Ti02 product.
The analytical results were as follows:
AI <0.01 wt.%
B <0.01 wt.%
Ba <0.01 wt.%
Be <0.01 wt.%
Ca <0.02 wt.%
Cd <0.01 wt.%
Co <0.01 wt.%
Cr <0.01 wt.%
Cu <0.01 wt.%
Fe <0.01 wt.%
K <1 wt.%
Mg <0.01 wt.%
Mn <0.01 wt.%
Mo <0.01 wt.%
Na <0.02 wt.%
Ni <0.01 wt.%
Pb <0.01 wt.%
Si <0.02 wt.%
V <0.01 wt.%
Zn <0.01 wt.%
The resulting precipitate of Ti02 was oven roasted at 1000°C to
drive
CA 02289967 1999-11-17
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off the hydrated water. The weight loss was 10%. This product obtained was
> 99.0% Ti02. The composition of the Ti02 is given in Table 3.
Table 3
Composition of Ti02
TBP/DEHPA Extraction
S02 Reduction/ppt'n
Element Weight
Aluminum AI203 < 0.01
Boron B < 0.01
Barium Ba < 0.01
Beryllium Be < 0.01
Calcium Ca0 < 0.01
Cadmium Cd < 0.01
Cobalt Co < 0.01
Chromium Cr < 0.01
Copper Cu < 0.01
Iron Fe203 < 0.01
Magnesium Mg0 < 0.01
Manganese Mn < 0.01
Molybdenum Mo < 0.01
Phosphorus P205 < 0.01
Potassium K20 < 0.01
Sodium Na20 < 0.05
Nickel Ni < 0.01
Lead Pb < 0.01
Silicon Si02 < 0.05
Titanium Ti02 > 99.8
Vanadium V < 0.01
Zinc Zn < 0.01
CA 02289967 1999-11-17
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To make a direct comparison to the test above, a sample of TBP
raffinate was pH adjusted with 50 weight percent sodium hydroxide to the
point of precipitating titanium dioxide. The addition of caustic was over a
five
and one-half hour period at an elevated temperature of 40°C. A solid
sample
obtained was thoroughly washed and dried.
The precipitation barren solution measured at a pH equal to <0Ø A
multi-element analysis of the hydrolyzed product showed metal contaminants
higher than achieved above with DEHPA present. Analysis of key elements
are as follows:
Chromium <0.01 wt.%
Iron <0.01 wt.%
Magnesium <0.01 wt.%
Manganese <0.01 wt.%
Aluminum <0.01 wt.%
Vanadium 0.021 wt.%
Zinc 0.011 wt.%
Calcination of the precipitate will increase the concentration of Ti02,
but, will also elevate the levels of zinc and vanadium. The use of TBP/DEHPA
allows extraction of vanadium and zinc. Controlled hydrolysis of titanium
dioxide at a low pH may result in the prevention of co-precipitation of
chromium.
In another test, an aliquot of TBP raffinate was tested to quantitatively
determine the precipitation efficiency of titanium from TBP raffinate using
S02
and TBP/DEHPA. The extraction of Vanadium was carried out at a phase
ratio equal to one, but at a lower contact temperature of 50°C. The
organic/aqueous contact interval was extended to 2 hours at this lower
temperature while maintaining the reducing atmosphere with S02 gas. After
disengaging both phases, the aqueous solution was separated from the
loaded organic. Again, sulphur dioxide gas was bubbled into the aqueous
solution while the temperature was raised to 90°C or higher. The 150-
160
CA 02289967 1999-11-17
-25-
milliliters of treated TBP raffinate was diluted with 300 milliliters of
distilled
water to initiate precipitation of the titanium dioxide. The dehydration
period to
form Ti02 was extended for 2.5 hours. The resulting white precipitate was mild
HCI washed followed by a distilled water wash and then oven dried at
100°C.
The analytical results of the 80 milliliters precipitation barren and the
multi-
element analysis of the white precipitate are follows:
Precipitation Washed
Barren Ti02
(80 Precipitate
ml)
AI 1570 ppm AI < 0.01 wt
Cr 1310 ppm Cr < 0.01 wt
Cu 4.8 ppm Cu < 0.01 wt
Fe 1990 ppm Fe < 0.01 wt
Mg 4590 ppm Mg < 0.01 wt
Mn 790 ppm Mn < 0.01 wt
Ni 320 ppm Ni < 0.01 wt
Mo N/A Mo < 0.01 wt
Si <50 ppm Si <0.05 wt
Na N/A Na <0.05 wt
Ti 5590 ppm Ti
58.0
wt
V V
1200 <
ppm 0.01
wt
Zn 21 ppm Zn < 0.01 wt
The analytical results above clearly show that the dehydration of
titanium dioxide takes place at a low pH (pH<0.0), and other metal
contaminants will not co-precipitate. The Ti02 product obtained is a hydrated
oxide that requires further roasting to produce a >99.8% pure solid. The
precipitation efficiency of titanium calculates to be 94.8% based on the assay
results above.
EXAMPLE VIII
A series of leaching tests were aimed at exploring the phenomenon
that occurred when leaching at about 90°C that caused some titanium in
solution to precipitate during HCI extraction of ilmenite. At high chloride
CA 02289967 1999-11-17
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concentrations and low pH, iron can form anionic chloro complexes and halo
metallic acid (H+FeCI-4) and remain in solution after leaching. Titanium forms
anionic chloro complexes at high chloride concentrations and low pH, H2TiC16
with relatively less efficiency. It is possible to dehydrate dissolved
titanium
compounds and obtain titanium dioxide compounds. This approach was
tested in a series of subsequent tests, to potentially eliminate additional
iron
reduction with S02 and liquid/solid separation stages. Another additional
advantage of this approach is the elimination of sulphate in leaching,
improving effluent recycling for pyrohydrolysis, potentially eliminating
sulphur
in the Ti02 product and hence in subsequent purification tests.
These tests used a concentrate that contained 19.7% Ti. There were
six tests performed. The first two tests produced 91.7% and 97.6% of the Ti in
the precipitate and 98.6% and 96.7% of the Fe in solution. Excess HCI was
added and the Ti in the precipitate was decreased. The leaching time was
decreased to 30 minutes from 2 hours and the amount of Ti in the precipitate
was decreased to 80%. In another test, the leaching temperature was held at
70°C for two hours then the temperature was increased to 95°C
for one hour.
The precipitate contained 53.1 % Ti.
A further set of leaching tests were aimed at exploring the
phenomenon that occurred when leaching at about 95°C that caused some
titanium to precipitate with the use of an oxidant. The oxidants tested were
oxygen gas and sodium chlorate (NaCl03). Also for these tests, the
precipitate was screened on 200 mesh and the two fractions were assayed for
Ti and Fe. Screening was used because as a precipitate was formed it would
be fine material as compared to unleached material (concentrate). For these
tests, the acid level was maintained at 44% excess and oxidant was added
only after 2 hours of leaching. These tests demonstrated that the Ti can be
precipitated with 95% of the Ti reporting to the minus 200 mesh fraction; the
iron level was about 1 %. The results were reproducible with a repeat test and
the amount of material treated was increased to 200 grams. The amount of
iron extracted and present in the leach solution was about 97% for most of the
tests.
CA 02289967 1999-11-17
-27-
Another set of tests was performed to optimize the amount of acid and
oxidant used. Lowering the acid used to stoichiometric amount and lower
amounts of oxidant caused the amount of Ti reporting to the fine fraction to
decrease to about 60% and iron level increase to 1.5 to 2% in this fine
fraction. If the oxidant was held constant at 1.5 g of NaCl03 per 100 grams of
concentrate and the acid was increased to 20% or 30% excess, the Ti
reporting to the fine fraction was 86% with 0.31 to 0.51 % iron at the
respective
acid excess level. On the basis of these tests, an optimum condition is 3g of
NaCl03 per 100 grams of concentrate and 20% excess acid, under which the
Ti reporting to the fine fraction is greater than 90% with an Fe level of 0.5%
in
the fine fraction. Oxygen was tested using the same conditions of acid, time
and temperature and 86% of the Ti reported to the fine fraction and the Fe
level was 0.9% in the fine fraction.
A multi-elemental analysis was conducted on the precipitation
products, and two results are presented in Table 4. Further details of these
tests are provided in Tables 5 and 6.
CA 02289967 1999-11-17
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Table 4
Composition of Ti02 Products made in the Testwork
Leach Test A Leach Test B
200 Mesh 200 Mesh
Leach Residue Leach Residue
Element Weight % Weight
Aluminum AI203 0.32 N/A
Boron B < 0.01 NA
Barium Ba < 0.01 N/A
Beryllium Be < 0.01 N/A
Calcium Ca0 0.16 1.56
Cadmium Cd < 0.01 N/A
Cobalt Co < 0.01 N/A
Chromium Cr < 0.01 N/A
Copper Cu < 0.01 N/A
I ron Fe203 0.96 0.803
Magnesium Mg0 0.33 N/A
Manganese Mn < 0.01 NA
Molybdenum Mo < 0.01 N/A
Phosphorus P205 N/A N/A
Potassium K20 N/A N/A
Sodium Na20 0.13 N/A
Nickel Ni < 0.01 N/A
Lead Pb < 0.01 N/A
Silicon Si02 3.74 N/A
Titanium Ti02 > 94.35 > 97.0
Vanadium V < 0.01 N/A
Zinc Zn < 0.01 N/A
CA 02289967 1999-11-17
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