United States Patent [19]
Miyoshi et al.
[11] 3,929,461
[45] Dec. 30, 1975
8 Claims, No Drawings
Primary Examiner-M. J. Andrews
Attorney, Agent, or Firm-Sheridan, Ross & Fields
[56] References Cited
UNITED STATES PATENTS
1,534,819 4/1 925 von Seth 75/60
3,753,681 8/1973 Vojkovic 75/24
Vanadium and titanium values are recovered from vanadium
containing titaniferous iron ore by subjecting
the ore to a reduction process to separate the iron
from the slag containing the vanadium and titanium,
fusing the slag with an alkali metal salt, oxidizing the
vanadium to its plus five oxidation state, and recovering
the, vanadium and titanium by conventional techniques.
[57] ABSTRACT
[52] U.S. CI 75/30; 75/24; 423/68
[51] Int. CI.2 C2lB 3/04
[58] Field of Search ; 75/21, 24, 30; 423/68
[54] FUSION-OXIDATION PROCESS FOR
RECOVERING VANADIUM AND TITANIUM
FROM IRON ORES
[75] Inventors: T. Kenneth Miyoshi, Lakewood;
Cornelius E. Berthold, Littleton;
Frank M. Stephens, Jr., Lakewood,
all of Colo.; Alfred K. Schellinger,
South Perth W.A., Australia
[73] Assignee: Ferrovanadium Corporation N.I.,
Perth, W.A., Australia
[22] Filed: Feb. 27, 1974
[21] Appl. No.: 446,156
3,929,461
2
SUMMARY OF THE INVENTION
The critical aspect of this process is the fusion of the
material containing the vanadium with an alkali metal
salt, followed by oxidation of the vanadium. It is be-
30 lievedthat the vanadium exists in the slag in the form of
spinel-type crystallites, and while in this form the vanadium
is not easily subject to the necessary oxidiation
reaction. By producing an intermediate fusion product
with an alkali metal compound the slag is physically
altered and the vanadium does not crystallize in the
same manner. This greatly facilitates the oxidation.
In order to produce the necessary molten slag at least
a critical amount of alkali metal salt must be added to
the vanadium-bearing material for a given temperature
40 prior to or during melting. Excess alkali metal salt can
be employed, but this is economically not advantageous.
The amount of alkali metal salt which must be
added to produce fusion and the resulting altered slag
at a given temperature varies for different ore samples,
and is dependent upon the impurities in the ore samples,
the amount of vanadium and other factors. Therefore,
prior to employing the process of the invention to
a particular ore or slag, experimental testing is required
to obtain the most desirable parameters. For most ores
SO and slags it has been found that for a temperature of
about I,400°C to I ,500°C an alkali metal compound in
the range of preferably about I to about 50 percent,
more preferably from about 2 to about 25 percent, and
most preferably from about 5 to about 15 percent
based on slag weight should be employed. However, it
must be emphasized that these ranges are provided
only as average values as one particular ore sample may
require,Jor example, 13 percent by weight of an alkali
metal compound, while another may require only, for
60 example 2percent.'
Similarly the temperature necessary to produce the
fusion reaction is a function of the amount of alkali
metal employed, as well as the amount of vanadiulTl in
the ore sample, the amount of impurities, etc. Therefore,
again experimentation must be performed on a
giveno~e sample in order to determine the optimum
conditions of temperature for that sample. Bearing in
mind that the critical aspect of the process is to pro-
.1
FUSION·OXIDATIONPROCESS FOR
RECOVERING VANADIUM AND TITANIUM FROM
. IRON ORES .
BACKGROUND OF THE INVENTION
Vanadium and titanium are recovered from slags and
ores .containing vanadium values by fusing the vana-
5 diumcontaining raw material with an alkali metal salt
and oxidizing the resulting fusion product either in the
1. .Field of the Invention . molten state or subsequent to solidification in order to
This process relates to the recovery of chemical com- oxidize the vanadium to its plus five oxidation state.
pounds, and more specifically to rare element com- The vanadium can then be recovered by conventional
pound recovery as classified in Class 23, subclass 15. 10 techniques, leaving the titanium-bearing residue sub-
2. The Prior Art: stantially vanadium free.
Numerous .vanadiumrecovery processes are dis- It is preferred to initially subject the vanadium-conclosed
in the prior art. It is well known; as for exampletaining ore to electric smelting under reducing condiin
U.S. Pat. No. 3,733,193, ·to recover vanadium by .Hons in·the presence of sufficient alkali metal salt so as
means of a: "salt roast" processwhereby an alkali metal 15 to melt the ore and reduce the contained iron to pig
salt is admixed with the vanadium-containing ore or iron. The vanadium and titanium values (plus impurislag
and the mixture roasted under oxidizing conditions ties such as silica and alumina) remain in the molten
at a suitable temperature. The resulting V2 Os is recov- slag phase, whi~e ~he i~on separates from thi~ slag.
ered by water or dilute acid leaching. While this pro- . The molten pIg Iron IS removed from the mIxture and
cess is of value, it is not usually possible to lower the V2 20 ~he molten sla~ is ?xidized to ~onvert ~he v.anadium to
o level in the leached residue to below I 00% to ItS plus five OXIdatIOn state. ThIS vanadIUm IS then con-
0.;5%. Such a relatively high V2 Os level inthe'leached vention~llyrecovered leaving a substantially vanadiumresidue
makes the leached residue unfit for some uses, free reSIdue.
i.e., for use in titanium pigment manufacture in cases 25 DESCRIPTION OF THE PREFERRED
where .theleached residue contains a suitably high EMBODIMENTS
titanium level.
The process described in U.S. Pat. No. 3,486,842 to
Michal discloses an improved vanadium recovery by
roasting the ore in a two~stage operation in the presence
of an alkali metal salt. This process necessitates
two separate roasts; the first designed to pelletize the
ore combined with the alkali metal salt and the second
serving as the actual roast. This obviously necessitates
addtitional expensive equipment over a single roast 35
process. Also this process utilizes a sintering technique
which results in relatively highly impure titanium values.
U.S~ Pat. No. 2,270;444 to Jenness discloses a process
for recovering vanadium by forming an intermediate
alkali earth metal salt fusion product with the ore.
Other similar references have disClosed such a fusion
technique, but operate under reducing .conditions
wherein the vanadium is recovered along with the iron.
The various prior art processes, exemplified by the 45
above referenced patents, each have drawbacks that
result in a low vanadium recovery along with leached
residues containing not less than 1.00 to 0.75% V2 Os,
when low grade vanadium-hearing ores or slags ( I% to
2% V2 Os) are used as the feed material. In contrast,
the process of the present invention can result in an
increased vanadium recovery' when operated on lowgrade
vanadium-bearing ores and slags; SUch that
leached residues of 0.20 to 0.40% V2 Os content are
obtained. This is important notonly .from a vanadium 55
recovery standpoint, but also from the standpoint of
utilizing the residue following the leach; for example, in
the manufacture oftitanium pigments. If this residue
contains too much vanadium, Le'.,O.75% or more V2 Os
content, the value of this residue for titanium pigment
manufacture is greatly reduced, as elaborate and expensive.
purification steps are required~p remove this
vanadium during the manufacture of such titanium
pigments.
The process of the present invention has been found 65
to produce high vanadium recoveries from relatively
low grade slags and ores and also yi~ld a leach residue
containing a low V2 Os level. .
3,929,461
EXAMPLE I
EXAMPLE 2
EXAMPLE 3
EXAMPLE 4
A slag, altered as in Example I andcontaining 1.96%
V2 0 5, was air oxidized at 750°C for I hour. Roasting
with a salt addition was .omitted from this test. Leaching
of the calcine was performed in the manner described
in Example I for leaching of the salt rO,ast calcine.
The final leach residue contained 0.74% V2 0 5 or
65.9% of the contained vanadium had been extracted'.
Vanadium extraction of 48.3% had been obtained by
4
tion and contribute impurities to t~e. final product.
Hence when acid leaching is employed the value. of
additional vanadium extracted must be balanced
against this disadvantage.
The vanadiumcbearing solutions obtained by leaching
can be processed by various means well known to
those skilled in the art. For example, vanadium may be
precipitated as ammonium metavanadate after concentration
and purification is accomplished in a' solvent
10 extraction system.
The preferred alkali metal compounds used in conjunction
with this process are the sodium salts, due to
their availability. Particularly sodium hydroxide, sodiumcarbonate,
sodium chloride and' sodium sulfate
are effective and easily obtainable. It must be understood
however that from a purely processing standpoint
compounds such as those of potassium, lithium, rubidium,
and cesium may also be used in practicing the
invention.
50 The original titanium slag of Example 1- was fused
with sodium sulfate equal to 10% of the slag weight.
The altered slagcontained 1.83% V2 0 5 after grinding
and removal of the magnetic particles larger than 100-
mesh. The salt roasting and leaching were performed as
in Example I. The final leach residue contained 0.46%
V2 0 5 or 74.6% ofthe vanadium had been extracted.
The water leach, prior to leaching with acid, had extracted
66.4% of the contained vanadium.
A titanium slag containing a vanadium equivalent of
1.77% V2 0 5 was fused at I400°C with sodium carbonate
equal to 10% of the slag weight fora period of 2
hours. Following air cooling of the melt and grinding,
the magnetic particles larger in size than 100-mesh
were removed, thereby increasing the vanadium .content
of the altered slag to 2.0% V205' The altered slag
was salt roasted wi~h sodium chloride equal to 10% of
30 the altered slag weight for two hours at 850°C in an
oxidizing atmosphere containing water vapor. The residue,
after acid leaching, contained 0.35% V205, equivalent
to 84.6% vanadium recovery. Vanadium.extraction
of 74.3% had been obtained by the water leach
prior to leaching with acid.
3
duce fusion or melting it becomes a matter of economic
priorities in selecting the proper balance between the
amount of alkali metal salt and the temperature. Average
values of temperatures are preferably from about
1300°c to about I700°C, more preferably from about 5
1400°c to about 1600°c and most preferably from
about I400°C to abolJt I500°C, but of course the temperature
must be at leastsufficient to produce fusion of
the slag with the alkali metal compound.
It is generally preferred that for a given ore body
containing both iron and vanadium that the iron be
recovered prior to vanadium recovery. This is accomplished
by means well known in the art, as for example
"reducing the ore to separate the iron from the other
. constituents, herein referred to as slag. This separation 15
is facilitated by the addition of alkali metal salt prior to
the reduction process. The alkali metal salt used is then
available for the fusion reaction with the slag, although
generally there is an insufficient amount and additional
alkali metal salt usually must be added to the slag to 20
produce the fusion. While it is generally preferred to
perform the process of the invention on the slag, some
ore samples are amenable to direct vanadium recovery
without prior removal of the iron utilizing this process.
The slag fusion is preferably accomplished in the 25
same operation as the iron separation. Alternatively a
two-stage heating operation may be employed wherein
iron. is separated during the initial step and the slag is
fused during a secondary step..
, While it is generally preferred to carry out the process
of this invention upon a molten slag, such as de'"
scribed above, it is also possible to utilize an ore, or
.concentrate derived therefrom, without prior reinoval
of the iron.
After the alkali metal salt has been added to the 35
molten slag it is preferred to fully oxidize the constituentsof
the slag by an oxygen lancing, wherein pure
oxygen gas is blown into the molten slag. Sufficient The original titanium slag of Example I was fused
heat is generated by the oxidation of the ferrous iron to with sodium chloride equal to 20% of the slag weight.
the ferric state to keep the slag molten. The oxidation 40 Removal of the magnetic particles larger in size than
ofthe slag by oxygen lancing insures that all of the 100-mesh, after grinding, increased the vanadium convanadium
values are fully oxidized to the plus five va- tent of the altered slag to 1.84% V205' Subsequent salt
lance state so that they can form soluble alkali metal roa~ting and leaching, as performed. in Example I,
vanadates which are recovered upon subsequent leach- resulted in a final residue containing 0.40% V2 0 5,
ing of the slag. While pure oxygen is preferred from a 45. representing a 79.4% recove~Y,. The water leach, prior
processing standpoint, oxygen mixed with inerts,· for to leaching with an acid, had extracted' 63.8% of the
example air, may also be employed. contained vanadium.
It is generally preferable to conduct the oxidation
roast in the presence of water vapor to facilitate complete
oxidation of the vanadium when the oxidation is
conducted on a solidified slag. The roasted calcines
may then be cooled prior to introduction into the
leaching stage or may be quenched directly from the
roast. Quenching may be preferable when the vanadium
collects in a glassy phase formed in the roast as 55
the vanadium is more soluble in this form than in a
crystalline state which forms upon slower cooling.
Leaching of the vanadium from the salt roast calcines
with water generally is preferred providing sufficient
extraction is obtainable. The water leach solution will 60
generally become basic due to the alkali remaining
after the roasting operation. Any iron compounds present
in the calcines remain with solids in the basic leach,
thus simplifying further processing. An acid leach may
also be used to extract additional vanadium values, or a 65
combination water/acid leach may be utilized. Leaching
with acid may dissolve any iron present in the calcine
in sufficient quantities so as to hinder later separa3,929,461
5
leaching with water pri~t to the acid leach.
ExAMPLE 5
By way of comparison the original tinmiulTIslag of
Example I was directly salt roasted under the same salt
roasting conditions used in Example I without alkali
metal salt fusion. The maximum vanadium extraction
obtained upon salt roasting of the original slag. at the
conditions cited in Example 1was 60.4% after It:aching
with acid. The· corresponding water ·Ieach extracted
39.1% of the contained vanadium. The final leach residue
contained 1.02% V2 Os, thus making it unsuitable
for Ti02 pigment manufacture.
6
EXAMPLE 9
A series of fusion or melting tests were carried out on
a slag containing 1.8% vanadium, calculated as V2 Os
5 to determine the optimum quantity of alkali metal salt
required to maximize vanadium extractability. The
various mixtures of the slag and sodium carbonate were
fused at 1400°c (2552°F) for 2 hours, followed by an
oxidizing roast at 850°C (I 562°F) with supplementary
10 alkali metal salt additions. The oxidized products were
then water and acid leached in the usual manner yielding
the following results:
EXAMPLE 6 15 QUANTITY OF
SODIUM CARBONATE
PERCENT VANADIUM
EXTRACTED
66%
70%
86.6%
2%
5%
10%
30
What is claimed is:
1. A process for the recovery of vanadium from
vanadium-bearing ores comprising:
reducing the ore to recover substantially all of the
25 iron and to produce a slag containing the vanadium,
treating the slag with an alkali metal salt in sufficient
quantity to produce a completely fused product
upon sufficient heating of the slag and so heating
the slag and alkali metal salt thereby forming an
altered slag,
oxidizing the molten altered slag in order to oxidize
the vanadium to its plus five valance state,
leaching the slag to extract the oxidized vanadium,
35 and
recovering the vanadium from the leach solution.
2. The process of claim 1 wherein an alkali metal salt
is introduced prior to the reduction of the ore.
40 3. The process of claim 1 wherein the alkali metal salt
is a sodium salt.
4. The process of claim 1 wherein the slag is treated
with at least one percent (I %) based on slag weight of
the alkali metal salt.
5. The process of claim 1 wherein the slag is heated
to a temperature of about I300°C to 1700°c.
6. In a process for the recovery of vanadium and
titanium from vanadium-bearing titaniferous ores
wherein the process comprises reducing the ore to
50 recover substantiaIly all of the iron and to produce a
slag containing the vanadium and titanium, oxidizing
the molten slag to oxidize the vanadium to its plus five
valance state, and recovering the oxidized vanadium
from the slag, the improvement comprising:
55 fusing the completely slag with an alkali metal compound
prior to oxidizing the slag.
7. A process for the recovery of vanadium and titanium
from vanadium-bearing titaniferous ores comprising:
reducing the ore in the presence of an alkali metal
salt to recover substantially all of the iron and to
produce a slag containing the vanadium and titanium;
heating the slag in the presence of an alkali metal salt
to produce a completely molten altered slag,
oxidizing the molten altered slag in order to oxidize
substantially all of the vanadium to its plus five
valance state, and
EXAMPLE 8
These tests were carried out with no fusion or melting 45
of the slag with an alkali metal salt. The mixture of slag
and alkali metal salt was heated to below the melting
point, but to a temperature where sintering took place.
Two portions of 200 grams each of an electric furnace
slag, containing 2.07% vanadium calculated as V2 Os
were heated to 13600 t (2480°F) with, respectively,
10% sodium carbonate and with 20% sodium chloride,
for a period of 2 hours. No melting or fusion of the
charge took place in either case, but the mixtures were
heavily sintered.
After the heating step, both charges were ground to
pass a 65 mesh screen and roasted 850°C (I 562°F) for
two hours to insure oxidation of the vanadium present.
The material was then subjected to a combination of 60
water and acid leaching to extract the vanadium values
therefrom.
A total of 43% of the vanadium was extracted from
the slag used in the sintering roast with sodium carbonate,
while 18% of the vanadium was extracted where 65
sodium chloride was used in the sintering roast. The
resulting titanium-bearing residues therefore were inadequate
for titanium pigment manufacture.
EXAMPLE 7
For comparison purposes the titanium-bearing slag of
Example 6 was air oxidized to remove residual carbon
and was then admixed with 10% by weight of sodium
carbonate and fused at I440°C in a graphite crucible to
maintain a reducing environment.
Upon subsequent leaching this slag it was found that
only 1.5% of the contained vanadium could be extracted
by combined water and dilute acid leaching,
thus illustrating the necessity for an oxidative environment
in practicing this invention.
A titanium-bearing slag, containing 2.09% vanadium,
calculated as V2 Os, prepared by electric smelting a
mixture oftitanomartite ore, coal and about 8% sodium
carbonate, based on the weight of the slag forming 20
constituents present, was air oxidized to remove residual
carbon and then mixed with 10% by weight of sodium
nitrate and fused at 1430°c in an alumina crucible
to simulate vigorous oxidation of the slag. Upon
leaching the melt in water it was found that 73% of the
contained vanadium was water extractable and that an
additional 17% was recovered with a dilute acid leach.
The final devanadated residue contained about 0.20%
V2 Os and was an exceIlent starting material for titanium
pigment manufacture.
3,929,461
7
recovering subst;mtiallyall of the oxidized vanadium
from said altered slag.
8. A process for the recovery of vanadium and titanium
from vanadium-bearing titaniferous ores compris- 5
iryg:
treating the ore with an alkali metal salt;
reducing the ore with the alkali metal salt to recover
. iron and to produce a slag containing the vanadium
and titanium; 10
15
20
25
30
8
treating the slag with additional alkali metal sal~ such
that the slag contains at least I% based on slag
weight of said alkali metal salt;
completely melting the slag and alkali metal salt to
form an altered slag;
oxidizing the molten altered slag in order to oxidize
the vanadium to.itsplus five valance state; and
leaching the slag to extract the oxidized vanadium
and to produce a titnaium bearing residue.
* * * * *
,,:ii
35
40
45
50
55
60
65
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60
65
t-f�Xfm0�(D�denHorzOCR'>References Cited
UNITED STATES PATENTS
7/10969 Litz , 23-15 W
7/1969 !Platzke et al. 23-15W
3/1959 Zimmerley et al. 23-18 X
7/1960 Zimmerley et al. 23-24
4/1966 Churchward , 23~15 W
211970 Ziegenbaly et al. 23-23 X
1/1971 Proter et al. 23-22
3,455,'677
3,45:8,277
2,876,065
2,945,743
3,244,475
3,495,934
3,558,268
23-23, 24 R, 51 R
7
(f) stripping the loaded agent of (e) with an alkali
metal hydroxide;
(g) extracting rhenium values from the strip solution
of (f) with pyridine or pyridine derivative; and
(h) recovering rhenium from the pyridine extractant 5
by distilling off the pyridine.
2. The process of claim 1 in which metal ion impurities
are removed from the strip solution of .(b) before
crystallizing ammonium tetramolybdate in (c).
3. The process of claim 1 in which the anion exchange 10
agent in (a) is a tertiary amine ion exchange resin and the
stripping solution of (b) is ammonium hydroxide.
4. A process for recovering molybdenum and rhenium
values from pregnant acid leach solutions containing these
values together with other metal impurities and derived 15
from dusts and flue gases resulting from roasting relatively
impure molybdenite concentrate, said process comprising:
(a) extracting molybdenum and rhenium values from
the pregnant acid solution with a liquid water in- 20
soluble amine ion exchange agent;
(b) stripping the molybdenum and rhenium values
from the exchange resin with ammonium hydroxide
solution to form a strip solution containing the molybdenum
as ammonium molybdate and the rhenium
as ammonium perrhenate;
(c) crystallizing the molybendum from the strip solution
in (b) as ammonium tetramolybdate by adjust