3,890,426 Method of treating alunite ore

Patent Number/Link:

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

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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METHOD OF TREATING ALUNITE ORE

FIELD OF THE INVENTION

3,890,426

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.

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.

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

3,890,426

* * * * *

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).

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