Primary Examiner-Herbert T. Carter
Attorney, Agent, or Firm-Van C. Wilks; Herbert M.
Hanegan; Stanley L. Tate
United States Patent (19)
Stevens et al.
[54) METHOD OF TREATING ALUNITE ORE
[75) Inventors: Douglas Stevens, Golden, Colo.;
Helge O. Forberg, Owensboro, Ky.;
Larry D. Jennings, Arvada, Colo.;
Frank M. Stephens, Jr., Lakewood,
Colo.; Francis J. Bowen, Golden,
Colo.
[73) Assignees: Southwire Company, Carrollton,
Ga.; National Steel Corporation,
Pittsburgh, Pa.; Earth Sciences,lnc.,
Golden, Colo.
[22) Filed: Mar. 21, 1974
[21) Appl. No.: 453,225
1,189,254
1,191,105
1,195,655
2,120,840
2,398,425
3,652,208
[57)
7/1916
7/1916
8/1916
6/1938
4/1946
3/1972
[II) 3,890,426
(45) June 17, 1975
Hershman et al... 423/120
Hershman 423/122
Chappell 423/131
McCullough 423/127
Haff 423/120
Burk et al... 423/127
ABSTRACT
[52] U.S. CI 423/127; 423/111; 423/118;
423/120; 423/122; 423/131; 423/629;
423/339; 423/567; 423/530
[51) Int. CI COIf 7/02; COif 7/06
[58) Field of Search 423/120, 118, Ill, 127,
423/131,629,122; 75/97 R, 101 R
[561 References Cited
UNITED STATES PATENTS
1,070.324 8/1913 Chappell 423/131
WATER OF H't'DRATION
This invention relates to a method for recovering aluminum
hydroxide from ore containing alunite by
roasting the Gre to remove the water of hydration,
leaching the roasted ore with a weak base to remove
potassium and sulfate, extracting the aluminum content
with a mixture of sodium hydroxide and potassium
hydroxide, and precipitating aluminum hydroxide
crystals from the solution.
21 Claims, 3 Drawing Figures
ALUNITE
ORE SOLIDS
LiqUID
SOLIDS
AL(OH)3
WATER OF HYDRATION
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OESILICATION ,'-
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PRECI PllATION
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ORE
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1
METHOD OF TREATING ALUNITE ORE
FIELD OF THE INVENTION
3,890,426
2
FIG. 3 is a diagramatic representation of an embodiment
of the present invention depicting an optional
method of silica removal.
DESCRIPTION OF THE PREFERRED
EMBODIMENTS
SUMMARY OF THE INVENTION
DESCRIPTION OF THE PRIOR ART
The present invention concerns a method of recover- 5
ing aluminum hydroxide from ore containing alunite by
calcination, leaching with a weak base, digestion with
a mixture of sodium hydroxide and potassium hydroxide
and a subsequent precipitation of aluminum hydroxide
by cooling and seeding the resultant solution.
Various techniques have been proposed for recovering
alumina from ore containing alunite. Of the various
techniques disclosed in the prior art the general
method involves treating alunite ore with concentrated
sulfuric acid followed by roasting or vice versa, with
SOa recovered as a bi-product and subsequently converted
into sulfuric acid and reused in the process. Aluminum
is retained in solution as a sulfate. Potash (K20)
is then added at pH of between 1 and 2 to precipitate
alum [K2SO.AI2(SO.h 18H20]. After precipitation the
alum is then roasted to disassociate the aluminum sulfate,
with the production of SOa and aluminum oxide
which is then recovered by crystallization. Ordinarily in
the prior art practioners have used much effort and expense
to eliminate potash. U.S. Pat. No. 1,948,887
(Saunders) is representative of prior art techniques.
U.S. Pat. No. 1,406,890 (Pederson) further discloses
the precipitation of "potash alum" by the addition of
potassium sulfate to an acidic leach solution. Loevenstein
in U.S. Pat. No. 2,958,580 teaches the recovery
of aluminum as aluminum sulfate by digesting aluminum
ore with sulfuric acid.
Although each of the aforementioned techniques
may be useful for the particular application referred to,
none of these conventional techniques however is suitable
for recovering aluminum hydroxide from low
grade aluminum ore containing alunite, which consists
of aluminum, potassium, sodium, sulfate and water.
Such ores being domestic to the United States in large
quantities offer a relatively untouched source of aluminum.
Referring to FIG. 1, which is a general diagramatic
flow sheet of an embodiment of this invention, ore containing
what is commonly known as alunite, which has
10 an approximate empirical formula of [K2AI6(OH lt2.
(S04)4) Na2AI6(OH)12(SO.). and/or combinations
thereof, is roasted to remove the water of hydration,
leached with a weak base, and the liquid and solid components
separated. The solid product of this separation
15 is then digested with a mixture of alkali metal hydroxides
and the liquid and solid components separated in
a second separation step. The liquid portion resulting
from the second separation is then seeded or heated to
remove silica by precipitating sodium aluminum sili-
20 cate. The remaining liquid is then cooled and/or seeded
to precipitate and recover aluminum hydroxide.
Advantageously the alunite ore is roasted in the
roasting step at a temperature of from about 400°C to
about 850°C, preferably the ore is roasted at a tempera-
25 ture of from about 500°C to about 650°C, in order to
effect removal of the water of hydration. Advantageously,
the roasting step is carried out at atmospheric
pressure and the preferred temperature is maintained
for from about one-half minute to about six hours. The
30 residence time within the roasting step varies greatly
depending upon the type equipment used.
In the leaching step the roasted ore is advantageously
l~ached with a base selected from the group consisting
of ammonium hydroxide and alkali metal hydroxides at
35 a pH of between about 8 and about 12 to dissolve sulfates
and alkali metals. Preferably the leaching step is
carried out at a temperature of up to about 100°C and
for a time of from about five minutes to about two
hours. Ammonium hydroxide is the most preferred
40 base for use in the leaching step, and the preferred concentration
is from about 12.5 to about 32 grams free
ammonia per liter of solution.
The liquid and solid components from the leaching
step are separated in the first separation step by con45
ventional means such as thickener tanks, filters, belt
The present invention concerns a method for recov- extractor filters, and the like.
ering aluminum hydroxide from ore containing alunite The solid content separated is then digested with a
by using a low temperature roast followed by leaching mixture of alkali metal hydroxides having a concentrawith
a weak base and digestion with a mixture of so- tion of up to about 300 grams per liter caustic exdium
hydroxide and potassium hydroxide. 50 pressed as Na2COa. Preferably the alkali metal hydro x-
One object of the present invention is to provide a ides are sodium hydroxide and potassium hydroxide.
novel method for economically extracting aluminum The mixture ratio can vary from about 20 percent to
hydroxide from ore containing alunite. about 100 percent sodium hydroxide, to about 80 per-
A further object of this invention is to provide a novel 55 cent to about 0 percent potassium hydroxide. Preferaand
economical method for separating aluminum hy- bly the mixture contains in excess of 50 percent sodium
droxide from ore containing alunite which consists of hydroxide. Advantageously the digestion conditions
aluminum, potassium, sodium, sulfates and water. are atmospheric pressure, a temperature of from about
These and other objects, features and advantages of 80°C to about 110°C, and a digestion time of from
the present invention will be apparent from the follow- 60 about five minutes to about two hours.
ing decription and the accompanying drawings. The digestion product is then separated in the second
separation step by conventional methods such as thickener
tanks, filters, and the like. Excess silica is then removed
from the separated liquid content by heating the
65 liquid and/or by seeding the liquid with sodium aluminum
silicates. Advantageously agitation is applied to
this liquid portion during the removal of excess silica.
If atmospheric pressure is used in the heating step a
BRIEF DESCRIPTION OF THE DRAWINGS
FIG. 1 is a general diagrammatic representation of an
embodiment of the present invention.
FIG. 2 is a diagrammatic representation of an embodiment
of the present invention depicting bi-product
recovery.
3,890,426
EXAMPLE NO. I
Fifty (50) grams of alunite calcine were mixed with
water containing 32 grams per liter free ammonia so
that the slurry contained 17 percent solids. The resultant
slurry was heated to from about 85°C to about
90°C and agitated for two hours, the slurry was then filtered
and the cake washed with a solution consisting of
20 grams per liter free ammonia in water, and with water.
Upon analysis of the cake 92.5 percent of the potassium
present before leaching was removed by the
ammonia leach and 93.5 percent of the sulfate present
prior to leaching was removed. Only one percent of the
Al20 3present before leaching was extracted during this
step.
143 grams of the ammonia leach residue were digested
in 340 ml of mixed caustic having a caustic concentration
of 220 grams per liter as Na2C03. The slurry
was boiled at a pressure of one atmosphere with mechanical
agitation for 60 minutes and filtered. Upon
analysis the filtrate was found to contain 74 grams per
liter AI20 3 and 1.16 grams per liter Si02. When compared
with the AI20 3and content of the starting materials
it was found that 88 percent of the AI20 a present
prior to the leach of Example No. I had been removed
in this caustic digestion step.
EXAMPLE NO.2
A quantity of leached alunite calcine was digested as
in Example No. I. Boiling temperatures were used to
insure maximum alumina digestion. A paddle stirer was
used to provide agitation. After digestion, the mixture
was filtered by suction.
In the case of the solution reported herein, the proportions
were 1200 milliliters of 250 grams per liter
NaOH and 600 grams of leached alunite calcine. Due
to test losses, only about 850 milliliters of liquor were
obtained. Enough demineralized water was added by
washing the filter cake to provide one liter of liquor. At
this point the solution contained 200 gm/l free caustic,
91 gm/l AI20 3 and 3.60 gm/l Si02 •
143 grams of the ammonia leach residue were digested
in 340 ml of mixed caustic having a caustic concentration
of 220 grams per liter as Na2C03' The slurry
was boiled at a pressure of one atmosphere with mechanical
agitation for 60 minutes and filtered. Upon
analysis the filtrate was found to contain 72.5 grams
per liter AI20 3 and 1.03 grams per liter Si02. When
compared with the AI20 3content of the starting materials
it was found that 86 percent of the AI20 3 present
prior to the leach of Example No. I had been removed
in this caustic digestion step.
4
seeded with aluminum hydroxide crystals during the
cooling step to accelerate the rate of precipitation and
control the particle size of crystalline aluminum hydroxide.
The liquid from the precipitation step of FIG. 2 (sodium
and potassium hydroxide) optionally may be recycled
and used in the digestion step. The solid content
of the precipitation step may be washed with water or
with a dilute acid.
The aluminum hydroxide product from the precipitation
step of FlG. 1 optionally may be calcined (heated)
to form alumina (AI20 a).
The following specific examples are intended to be
illustrative of the invention, but not limiting of the
15 scope thereof.
3
temperature of about 90°C for a time of at least one
hour is required. If pressure in excess of atmospheric
pressure is used a temperature of from about 90°C to
about 200°C is required to precipitate the sodium aluminum
silicate in a time of at least 15 minutes. Advan- 5
tageously the heating is carried out at a pressure of
from about 0.5 atmosphere to about 7 atmospheres.
Aluminum hydroxide seed crystals may then be added
to the solution and upon cooling crystals of aluminum
hydroxide are formed, precipitated and are separated 10
from the solution as crystalline aluminum hydroxide.
Prior to the roasting step the alunite ore optionally
may be crushed to a particle size having a greatest distance
between parallel opposite exterior surfaces of
about one inch or less. Optionally the product may be
ground to a particle size of about eight mesh or less
subsequent to the roasting step.
Referring to FlG. 2 in more detail, the liquid from the
first separation step optionally may be processed by
vacuum or cooling crystallization to precipitate a mix- 20
ture of ammonium sulfate and potassium sulfate when
ammonia is the weak base employed in the leaching
step, or sodium and potassium sulfate when sodium hydroxide
and potassium hydroxide are the base. When
using ammonia, the preferred base, the mixture of am- 25
monium sulfate and potassium sulfate is removed from
the solution by filtering, centrifuging or the like. The
mixed salts can either be marketed as such or fed to the
pyrolysis unit shown in FIG. 2, where the ammonium
sulfate is pyrolyzed at a temperature of about 300°C to 30
about 400°C to yield ammonia, water, and sulfur trioxide.
The pyrolysis unit can be a fluidized bed reactor,
a rotating kiln, a shaft furnace or the like. Vapors from
the pyrolysis unit are then passed through a column of
pebble lime which reacts with the sulfur trioxide pro- 35
duced by the pyrolysis to form calcium sulfate. The ammonia
and water produced by the pyrolysis are also
passed through the lime column before being recycled
to the weak base leaching step. Calcium sulfate so produced
can then be either prepared for marketing or dis- 40
carded as a waste.
The liquid content separated in the first separation
step of FIG. 2 optionally may be processed by adding
a weak base, such as ammonia, thereby precipitating
potassium sulfate. The liquid may then be boiled in a 45
lime boil step in the presence of lime [Ca(OHhl. preferably
in excess of stoichiometric amounts at atmospheric
pressure, a reaction time of from about fifteen
minutes to about one and one-half hours. The product
of the lime boil step is then separated by conventional 50
means such as centrifuge, filter, thickener tanks, vacuum
distillation or crystallization, and the like. The liquid
portion then can be recycled to use in the leaching
step and the solid precipitated sulfate converted to 55
commercial products such as sulfuric acid, elemental
sulfur and the like.
Referring to FIG. 3 in more detail, the product
formed in the silica removal step optionally may be filtered
and the liquid solution containing aluminum hy- 60
droxide transferred to the precipitation step. The solid
content filtered is sodium aluminum silicate with or
without a sulfate ion depending upon the concentration
of silicon and sulfate in the solution.
After removal of silica (precipitated as sodium alumi- 65
num silicate) the resultant liquid is cooled to precipitate
crystalline aluminum hydroxide, which is then separated
from the liquid. Advantageously the liquid is
5
3,890,426
6
* * * * *
This invention has been described in detail with par- 9. The method of claim 1 wherein the alkali metal hyticular
reference to preferred embodiments thereof, it droxides of Step (d) are selected from the group conshould
be understood that variations and modifications sisting of sodium hydroxide and potassium hydroxide.
can be effected within the spirit and scope of the inven- 10. The method of claim 1 in which the precipitation
tion as described hereinbefore and as defined in the ap- 5 of silica of Step (f) is performed by heating the liquid
pended claims. to a temperature of about 90°C for at least one hour at
What is claimed is: atmospheric pressure.
1. A method for recovering aluminum hydroxide 11. The method of claim 1 in which the precipitation
from ore containing alunite comprising the steps of: of silica of Step (0 is performed by heating the liquid
a. roasting the ore to remove the water of hydration, 10 at a pressure of from about 0.5 atmosphere to about 7
b. leaching the roasted ore from Step (a) with a weak atmospheres at a temperature of from about 90°C to
base at a pH of from about 8 to about 12 to dissolve about 200°C and for at least fifteen minutes.
sulfate and alkali metals, 12. The method of claim 1 in which the precipitation
c. separating the liquid and solid portions of the of silica in Step (f) is accelerated by seeding with soslurry
resulting from Step (b), said liquid portion 15 dium aluminum silicates.
containing dissolved sulfate and alkali metals, 13. The method of claim 1 in which the precipitation
d. digesting the solid portion from Step (c) with an of aluminum hydroxide in Step (h) is performed by
aqueous mixture of alkali metal hydroxides at a cooling the liquid until crystalline aluminum hydroxide
concentration and at a temperature sufficient to is formed.
extract the aluminum content from said solid por- 20 14. The method of claim 1 further including accelertion,
ating the precipitation of aluminum hydroxide in Step
e. separating the liquid and solid portions of the (h) by seeding the liquid with aluminum hydroxide
slurry resulting from Step (d), crystals.
f. precipitating silica from the liquid portion resulting IS. The method of claim 1 including the additional
from Step (e), 25 step of washing the precipitation product of Step (i)
g. separating the liquid and solid portions resulting with water.
from Step (f), 16. The method of claim 1 including the additional
h. precipitating aluminum hydroxide from the liquid step of washing the precipitation product of Step (i)
portion resulting from Step (g), with an acid having a pH of about 4.5.
i. separating the aluminum hydroxide precipitate 30 17. The method of claim 1 including the additional
from the liquid portion resulting from Step (h). step of calcining the aluminum hydroxide precipitation
2. The method of claim 1 wherein Step (a) is per- product of Step (i).
formed at a temperature of from about 400°C to about 18. The method of claim 1 including the additional
850°C. step of crushing the ore containing alunite to a particle
3. The method of claim 1 wherein Step (a) is per- 35 size having a greatest distance between parallel oppoformed
at a temperature of from about 500°C to about site exterior surfaces of about one inch or less prior to
650°C. Step (a).
4. The method of claim 1 wherein the weak base of 19. The method of claim 1 including the additional
Step (b) is selected from the group consisting of ammo- step of reducing the size of the product of Step (a) to
nium hydroxide and alkali metal hydroxides. 40 a particle size of about 8 mesh or less before proceed-
S. The method of claim 1 in which Step (b) is per- ing to Step (b).
formed at a temperature of from about 20°C to about 20. The method of claim 1 including the additional
120°C and for a time of at least five minutes. step of recovering Si02 from the solid content sepa-
6. The method of claim 1 in which the sulfate sepa- rated in Step (e).
rated in Step (c) is converted to sulfuric acid. 45 21. The method of claim 1 including the additional
7. The method of claim 1 in which the sulfate sepa- step of filtering the solution formed in Step (f) to yield
rated in Step (c) is converted to elemental sulfur. sodium aluminum silicate solids and sodium aluminum
8. The method of claim 1 in which potassium sulfate sulfate solids.
is recovered from the liquid content of Step (c).
50
55
60
65
ize:��btf��0�y:"Times New Roman","serif";mso-fareast-font-family: HiddenHorzOCR'>second oxidation zone.
* >I< >I< * *
About 7% of the molybdenum contained in the calcines
is also solubilized in the sulfurous acid leach.
The leached residue is separated from the leach solution
by filtration and after drying is ready for packaging 5
for sale. The leach solution joins the solutions from the
scrubbers on the flash roaster and re-roaster.
The effectiveness of the above-described process is
graphically illustrated by the high recovery of rhenium
and molybdenum achieved. it provides for the recovery 10
of up to 95% of rhenium and high recovery ofmolybdenum
in molybdenite with a minimum of process time
and a minimum of oxygen and added heat. The economic
advantages of these features are apparent. The
process is adaptable to either a batch or continuous op- is
eration.
It is an attractive side advantage of the. process that
a small volume of exhaust gas containing a high percentage
by volume of sulfur dioxide is produced. The
process is normally operated with an exhaust gas volume
discharge rate of 1,350 cubic feet per minute
(CFM) with up to 220% excess oxygen and 30-50% by
volume of sulfur dioxide in the exhaust gas. This high
volume percentage of sulfur dioxide makes its recovery
economically feasible for various commercial uses. In
contrast, present-day processes utilizing air for cooling
and for supplying oxygen are of necessity operated with
an exhaust volume discharge rate of 40,000 CFM, 16
volume percent excess oxygen and 1-2 volume percent
of sulfur dioxide. This volume percentage of sulfur dioxide
in the exhaust gas is so low that its recovery is not
economically feasible because it involves processing
such large volumes of gas. As a result the sulfur dioxide
is exhausted to the atmosphere creating a serious pollution
problem in heavily populated areas. The process of 35
this invention eliminates this problem.
The reduced volume of exhaust gas also results in a
much higher concentration of rhenium oxide in the exhaust
gas than is obtained in conventional processes.
As a result, recovery of substantially all of the rhenium
is far more feasible and economical than in present processes
using air with resultant large volumes of exhaust
gas to be processed for recovery of the rhenium oxide.
Reduction of the volume of gas processed through
the system by a factor of about 30resultsin a dr1!§jic: 45
reduction in the size of equipment require-d~ith~jgnificant
savings in equipment cost and floor space.
What is claimed is:
n. A method for recovering rhenium and molybdic
mdde from molybdenite concentrate which comprises:
a. pre-heating particles of said concentrate in an oxygen-
free atmosphere to a temperature not in excess
of about 750"C to raise the temperature of the particles
to promote flash oxidation of the molybdenite
when the particles are introduced into a flash
oxidation zone,
b. causing said pre-heated particles to fall through a
first oxidizing zone of heated oxygen with said particles
and heated oxygen moving countercurrent to
each other to disperse said pre-heated molybdenite
particles in said heated oxygen to provide maximum
particle surface contact with heated oxygen
for effective oxidation, said first oxidation zone
being heated substantially by the exothermic heat
of the reactions occurring in said first oxidation 65