4,618,480 Recovery of alumina values from alunite ore

Patent Number/Link:

United States Patent [19J

Hazen et ale

[l1J Patent Number:

[45J Date of Patent:

4,618,480

Oct. 21, 1986

OTHER PUBLICAnONS

Kinetics of Regeneration of Spent Seed from MHO

Power Generation Systems, by J. I. Joubert, P. F. Mossbauer,

T. C. Ruppel, D. Bienstock, U.S. Energy Research

& Development Administration, Pittsburgh Energy

Research Center, Energy Conversion, Pittsburgh,

PA.

Engineering Design for the Westinghouse MHO Seed

Regeneration Process, 7th International Conference on

MHO Electrical Power Generation, vol. 1, pp.

351-355, by T. J. Lahoda and E. E. Lippert.

SCA-Billerud Recovery Process Goes On-Stream, by

E. Horntvedt, Pulp and Paper International, Aug. 1968.

Study of Seed Reprocessing Systems for Coal Fired

Open Cycle Coal Fired MHO Power Plants, Task I, 37 Claims, 4 Drawing Figures

A novel process for the recovery of alumina and potassium

sulfate from alunite is provided comprising leaching

the alunite with potassium hydroxide to which no

sodium materials have been added, said leach solution

being saturated with potassium sulfate. Aluminum values

are solubilized into the leachate, and potassium and

sulfur values are rendered soluble, but remain in the

residue. The leachate is desilicated if necessary, preferably

with lime, and aluminum trihydroxide is precipitated

therefrom, followed by calcining to alumina product.

The residue is leached to solubilize potassium sulfate

in a secondary leach and the potassium sulfate product

crystallized therefrom. Potassium hydroxide is regenerated

from a portion of the potassium sulfate secondary

leachate by several methods.

Novel procedures for regenerating alkali metal hydroxides

from the corresponding sulfates are also provided

including routes involving formates and carbonates as

intermediates and pyrohydrolysis.

Selection of Processes for More Detailed Study, DOE

Contract No. DE-AC02-79ET 15613, JuI. 17, 1980.

G. Hohorst and H. L. Hou, Chern. Abstracts, 6828E

(China).

Japanese Pat. No. 76-20, 438 (1973), Chern. Abstracts

86:191990F.

I. Gruncharov, Chern. Abstracts, 98-218029M (Bulgaria).

Comparative Characteristics of Sodium and Potassium

Hydroaluminosilicates Formed Under Conditions of

Silica Removal From Aluminate Solutions, by T. I.

Avdeeva and A. A. Novolodskaya, Journal of Applied

Chemistry of USSR (Engl. Trans.), vol. 39, No.2, pp.

271-277, Feb. 1966.

Recovery of Sodium-Base Pulping Chemicals by Bicarbonation

and Crystallization by Johna Gullichsen, Erik

Saiha and E. Norman Westerberg, Tappi, Sep. 1968,

vol. 51, No.9, pp. 395--400.

Two-Stage Disilification of Pure Potassium Aluminate

Solutions, at Atmospheric Pressure, by A. I. Lanier and

Mai-Ki, Soviet Journal of Non-Ferrous Metals, vol. 37,

No.9, pp. 55-57, 1964 (Engl. trans.).

Primary Examiner-H. T. Carter

Attorney, Agent, or Firm-Sheridan, Ross & McIntosh

[57J ABSTRACT

RECOVERY OF ALUMINA VALUES FROM

ALUNITE ORE

Inventors: Wayne W. Hazen, Denver; David L.

Thompson; James E. Reynolds, both

of Golden; Nicholas J. Lombardo,

Boulder; Paul B. Queneau; John P.

Hager, both of Golden, all of Colo.

Assignee: Resource Technology Associates,

Boulder, Colo.

Appl. No.: 641,020

Filed: Aug. 15, 1984

Int. CI.4 COIF 7/06

U.S. CI. 423/127; 423/120;

423/122; 423/183

Field of Search 423/120, 122, 127

References Cited

U.S. PATENT DOCUMENTS

3,134,639 5/1964 Nylander 23/63

3,890,425 6/1975 Stevens et al. 423/127

3,890,426 6/1975 Stevens et al. 423/127

3,983,211 9/1976 Nasyrov et al. 423/128

3,984,521 10/1976 Nasyrov et al. 423/120

3,996,334 12/1976 Hartman et al. 423/127

4,029,737 6/1977 Stevens et al. 423/127

4,057,611 1111977 Jennings et aI. 423/127

4,064,217 12/1977 Hartman et al. 423/120

4,230,678 10/1980 Hartman et al. , 423/112

4,331,636 5/1982 Svoronos 423/126

FOREIGN PATENT DOCUMENTS

590158 1211933 Fed. Rep. of Germany.

791021 9/1935 France.

[54J

[75J

[73]

[52]

u.s. Patent Oct 21, 1986 Sheet 1 of4 4,618,480

ORE FIGURE I

PREPARATION

20 'r

24 25

CaCO CaCO

MAKEUP~

LIMESTONE

PRIMARY 2' 12

AI 2 03 3

--4 10 :

DSP II

LEACH RECOVERY

(13) RESlo,

I : 26 1 127

I I 9

I -¥_.; IMPURITIES 28

-t._ ..; CONTROL ~IMPURITIES TO

DISPOSAL

15 SECONDARY 4

I _~

I TO DISPOSALI..-LE"TA_C_H~_....I

I \ 14

*I ------3-0 17 K2C03 6 22 ~~---+--~ t--~CaS04'2H20 BI If :::< ;<J GENERATION TO DISPOSAL :~ __ ~ f-~; ,16 ~ ~ 21

;- AJ K2C03

IK2S04 51 S. ~ o :Ti

REeOV!"' I Z 0 J

• ~ ,AUSTlClZATION

PRODUCT K2S04

l' -I

LIM E 8

REGENERATION

u.s. Patent Oct 21, 1986 Sheet 2 of4 4,618,480

,...., a

:I:

CL.. ......

~ -I

K2SCS)

-2

FIGURE 2

PHASE STABILITY DIAGRAM

K-O-H-C-S SYSTEM

K2C03(S)

-I 0 I 2

LOG (PC02/PCO)

TEMPERATURE = 700 DEG. C

. K2S04( S)

u.s. Patent Oct 21,1986 Sheet 3 of4

FIGURE 3

4,618,480

STABILITY DIAGRAM FOR THE K-O-H-C-S SYSTEM AT 15% CO 2 ; PT =I ATM

~ 0

~ K2SCS.L)

N -I

u.s. Patent Oct 21, 1986 Sheet 4 of4

FIGURE 4

4,618,480

ALUMINUM EXTRACTION vs TIME

zo;:::

X""""

IJ.I

100' C

CONDITIONS:

START A/!<,Y : 0,23

TARGET A/K.!I : 0.60

FINAL KOH, -1909/1

RECOVERY OF ALUMINA VALUES FROM

ALUNITE ORE

This invention is in the field of hydrometallurgy, and

particularly relates to a process for the selective recovery

of aluminum, potassium and sulfur values from alunite

ore using a potassium hydroxide leach and to methods

of producing alkali hydroxides, such as potassium

hydroxide from alkali sulfates.

4,618,480

TECHNICAL FIELD

French Pat. No. 791,021 teaches a process for leaching

alunite with a KOH leach and solubilizing the aluminum,

potassium, and sulfur values from the ore into

the leachate. Potassium, sulfur and silicate values are

5 crystallized from the leachate by cooling, with subsequent

processing of the leachate to recover aluminum

values. The process of French Pat. No. 791,021 is directed

primarily toward production of pure alumina

and does not teach any overall system demonstrating

10 recovery of pure K2S040r regeneration ofKOH by any

of the methods utilized herein.

Other references to potassium hydroxide leaching of

BACKGROUND OF THE INVENTION alunite include G. Hohorst et al.• J. Kim Enge. (China), 4,

Alunite is a potassium aluminum sulfate mineral hav- 15 21-8 (1937); Chemical Abstracts-6828E; Japanese Pat.

ing the general formula: KAI3(OHMS04h. Alunite No. 76-20,438 (1973); Chemical Abstracts, 86, 191990F,

ores, also typically contain varying amounts of sodium- and I. Gruncharov, Chemical Abstracts 98-218029M

containing minerals and/or silica, Si02 Several pro- (Bulgarian).

cesses have been developed for recovering aluminum None ofthis prior art discloses or suggests the separavalues

from alunite ore, many of which also include 20 tion of aluminum values into a primary leach liquor by

recovery of the potassium values as K2S04. All have using a potassium hydroxide leach saturated with

been plagued by some major economic flaw; for exam- K2S04 and thereby leaving the potassium and sulfur

pie, expensive purchased reagents, complicated and values in the primary leach residue.

capital-intensive processes, requirements for sulfuric J. Gullichsen et al., "Recovery of Sodium-Base Pulpacid

production to handle sulfur dioxide off-gas, efflu- 25 ing Chemicals by Bicarbonation and Crystallization,"

ent pollution problems, and/or high energy require- Tappi, Vol. 51, No.9, 395-400 (Sept. 1968) and U.S.

ments for thermal pretreatment of the ore. Pat. No.3, 134,639 disclose sulfidization and carboniza-

Alunite is not very soluble in water and, as such, tion reactions for converting alkali metal sulfates to

many mineral recovery processes involve caustic leach- carbonates, but this art does not show subsequent coning

of the ore to solubilize the potassium and aluminum 30 version to the hydroxide nor recycle of the hydroxide

values into the leach liquor from which they are subse- solution to a leach process for alunite. Similarly, Gerquently

separated and recovered. U.S. Pat. Nos. man Pat. No. 590158 (1933) shows conversion ofpotas-

3,983,211; 3,890,425; 3,890,426; 4,029,737; 4,064,217 and sium sulfate to potassium formate with subsequent con-

4,057,611 exemplify prior art teachings with respect to version to potassium carbonate, but this art does not

caustic leaching with NH40H or NaOH, KOH and 35 show or suggest subsequent conversion to the hy'droxmixtures

thereof. Many prior art processes such as those ide, nor recycle of the hydroxide to an alunite leach

of U.S. Pat. Nos. 3,890,425; 3,890,426; 4,029,737 and process.

4,057,611 require roasting or dehydrating the alunite The pyrohydrolysis of sodium sulfate to sodium carprior

to leaching. In each of these prior art processes, bonate in a green processing with coal and water has

the potassium values of the alunite ore are extracted in 40' been described in E. Horntvedt, "SCA-Billerud Recovthe

initial leaching consistent with what has heretofore ery Process Goes On-Stream," Pulp and Paper Internabeen

considered the only effective method of mineral tional, August, 1968. Pyrolysis of potassium sulfate with

value recovery. In such methods, the alumina-contain- coal or reducing gases has also been described in E. J.

ing residue is treated in a typical Bayer-type circuit for Lahoda et aI., "Engineering Design for the Westingrecovery

of high grade alumina, i.e. leached into solu- 45 house MHD Seed Regeneration Process," 7th Internation

typically a caustic solution of NaOH and a mixture tional Conference on MHD Electrical Power Generaof

NaOH and KOH, and reprecipitated. tion, Vol. 1, Page 351; J. I. Joubert et al., "Kinetics of

It is known that where significant amounts of silica Regeneration of Spent Seed from MHD Power Generaare

present in the ore, losses of aluminum values by tion Systems," U.S. Energy Research and Development

precipitation of insoluble alkali aluminosilicates can 50 Administration, Pittsburgh Energy Research Center,

occur. U.S. Pat. Nos. 3,983,211 and 3,984,521 teach Energy Conversion, Pittsburgh, Pennsylvania; and

leaching processes wherein production of aluminosili- "Study of Seed Reprocessing Systems for Coal Fired

cates is minimized. In both patents, a mixture of KOH Open Cycle Coal Fired MHD Power Plants, Task I,

and NaOH is used as the initial leach. In addition, U.S. Selection of Processes for More Detailed Study,"

Pat. No. 3,984,521 teaches that the initial leach must be 55 D.O.E. Contract No. DE-AC 02-79ET 15613, July 17,

carried out at a temperature below 60° C. with the 1980. These authors all require a two-stage reaction in

disadvantage of slow kinetics. U.S. Pat. No. 3,983,211 which potassium sulfate is converted to potassium sulpermits

a higher temperature leach but requires a sub- fide at high temperature, followed by low temperature

stantial excess of sodium ions relative to potassium ions. oxidation of the potassium sulfide to potassium carbon-

In U.S. Pat. No. 3,983,211 the aluminum and potassium 60 ate, and none teaches or suggests the parameters or

ore values are both extracted into the initial leach li- viability of a process for the direct, one-stage pyrolysis

quor. U.S. Pat. No. 3,984,521 teaches solubilizing the conversion of an alkali metal sulfate to the correspondaluminum

values in the relatively low temperature ing carbonate. In addition, none teach or suggest further

leach while leaving potassium and sodium sulfates and conversion of the carbonate to the hydroxide.

silicates in the leach residue, but because of the substan- 65 None of the above-described art shows or suggests

tial amount of sodium present requires a separation the desirability of using alkali metal and sulfur values

between the sodium and potassium sulfates which are contained in an ore to generate a caustic leach solution

subsequently extracted together from the residue. for that ore.

4,618,480

SUMMARY OF THE INVENTION

the crystals calcined to oxidize the formate to the corresponding

carbonate. The carbonate is then causticized

This invention provides novel processes for recover- as described above, preferably followed by decomposiing

aluminum, sulfur and potassium values from alunite tion of the formed carbonate to lime for recycle to

ore by contacting alunite ore in a primary leach with an 5 causticization.

aqueous potassium hydroxide solution which is satu- Alternatively, in a novel improvement on processes

rated with potassium sulfate and to which no sodium incorporating the formate regeneration procedure, the

has been added in the form of sodium hydroxide or conversion from the sulfates to the hydroxides is a diotherwise.

By this potassium hydroxide leach the alumi- rect hydrometallurgical one. The potassium sulfate is

num values are selectively removed into the leach solu- 10 digested under pressure with excess lime and carbon

tion while rendering the potassium and sulfate values in monoxide, followed by the addition of an oxidizing

the ore water soluble but nevertheless left in the residue agent to the reaction vessel to form carbonate and then

of the primary leach. After separation of the primary potassium hydroxide. This conversion process is an

leachate from the residue, the leachate is optionally advantageous improvement over the known "formate"-

desilicated and then treated to precipitate aluminum 15 route conversion of alkali metal sulfates to their hydroxtrihydrate

from the leach solution. The aluminum trihy- ides, independent of its use in conjunction with the

drate is calcined to produce an alumina product. The alunite value recovery processes of the present invenpotassium

values are recovered from the residue for tion. Calcium carbonate by-products may be decomultimate

production of potassium sulfate and, in a pre- posed to form lime for recycle to the process.

ferred embodiment for regeneration of the KOH 20 In a separate embodiment, the subject alunite processneeded

for the primary leach.

The potassium sulfate-containing solid residue from ing is coupled with KOH regeneration by pyrohydrolth

e pn.mary 1each' 1 h d . d 1 h 'th ysis. A novel one stage potassium hydroxide generation IS eac e m a secon ary eac WI by pyrohydrolysis has been disovered wherein sulfates

an aqueous solution, preferably water and/or recycle

solution, i.e. mother liquor returning from the potas- 25 are converted to the corresponding carbonates by reactsium

sulfate crystallizer circuit, to solubilize the potas- ing the sulfates with a reducing agent at elevated temsium

sulfate in the alunite residue. After separation of peratures for a time sufficient to produce solid carbonthe

leachate containing the solubilized potassium sulfate ates which may advantageously be used independent of

from the solid residue, the potassium sulfate is recov- whether the alkali metal sulfate emanates from the aluered

by crystallization. 30 nite processing of the present invention. When incorpo-

Several embodiments of the present invention utilize rated with the novel alunite leaching processes dedifferent

hydroxide generation schemes wherein the scribed herein, the carbonates are then causticized as

potassium sulfate (K2S04) produced during processing above described, preferably with decomposition of the

is in part used to regenerate the potassium hydroxide formed calcium carbonate to lime and carbon dioxide

(KOH) for recycle to the initial leaching step. In one 35 and recycle of the lime to causticization. This proceinstance,

a novel method of converting K2S04 to KOH dure has the advantage of producing hydrogen sulfide

by pyrohydrolysis is provided which can be used with from which by-product sulfur can be recovered.

or without the alunite leaching processes of the present A further alkali metal hydroxide generation process

invention. In another instance, a novel process of pro- useful in conjunction with the alunite leaching process

ducing KOH from K2S04 through use of a formate 40 of the present invention involves reacting the sulfates

intermediary is disclosed which can also be used other with barium oxide (or hydroxide) to form the potassium

than in conjunction with the alunite leaching processes hydroxides directly, along with barium sulfate. This

of the present invention. barium sulfate may then be used to regenerate barium

The first of these combined leaching leach regenera- hydroxide by reducing with coal, e.g. in a black ash

tion processes involves the use of the Al and K value 45 rotary furnace, to barium sulfide, then reacting with

recovery process of the present invention coupled with carbon dioxide to form barium carbonate and hydrogen

use of the known Nylander process (U.S. Pat. No. sulfide, followed by reduction, using coal or coke as a

3,134,639) for the conversion of alkali metal sulfates to reductant, of the barium carbonate to barium oxide for

their corresponding carbonates. Formation of carbon- recycle to the alkali-metal-hydroxide-forming reaction.

ates is followed by causticization of the formed carbon- 50 In preferred embodiments of this invention, the levels

ates with lime to generate: (1) hydroxides for recycle to of impurities in the alunite processing steps are conthe

primary leach; (2) calcium carbonate which is de- trolled by treating bleed streams from the leaching solucomposed

to carbon dioxide for recycle to the Nylander tions. The treated products may then be returned to the

reactions and lime for recycle to the causticization step. leach circuits.

This method has the unexpected advantage of a bal- 55 As will be known and understood by those skilled in

anced series of process steps wherein by a unique series the art, the processes described above for converting

of recycle streams, AI, S, and K values are recovered in potassium sulfates topotassium hydroxide for recycle to

a total process wherein the only consumed reagent, if the KOH leach step may be modified and incorporated

desired, can be CaO. Optionally, this CaO can be pro- for use in any caustic leaching process for ores containduced,

e.g. in a kiln, from CaC03 which would then be 60 ing alkali metal and sulfur values in which alkali metal

the only chemical reagent purchased. sulfate-containing streams may be generated to the re-

A different embodiment couples the novel alunite quired caustic. Examples of such process are those of

leaching process ofthe present invention with a hydrox- U.S. Pat. Nos. 3,984,521; 3,983,211; 4,230,678;

ide generation step comprising digestion under pressure 3,890,425; 3,890,426; 4,064,217; and 4,029,737 incorpoof

the corresponding potassium sulfates with lime and 65 rated herein by reference. As such, other embodiments

carbon monoxide to form the alkali metal formates and of this invention comprise novel processes which are

gypsum. The gypsum is removed and the clear liquor improvements over known processes for caustic leachcontaining

the alkali metal formate is evaporated and ing of ore in that they incorporate a particular leach

4,618,480

minImIZing precipitation of the aluminum values as

aluminosilicates.

Aluminum trihydroxide is precipitated from the first

leachate, preferably after it has been desilicated, as silica

5 is an unacceptable contaminant in alumina for most

purposes, e.g. electrolytic reduction. Desilication methods

known to the art may be used; and in a preferred

embodiment of this invention involving conducting the

process so as to use lime as the only makeup reagent,

10 desilication is accomplished by adding lime to the leachate

to precipitate the solubilized silica as calcium aluminosilicate.

In the context of this invention, "lime,"

"CaO" and "Ca(OH)2" are used interchangeably, it

being understood that CaO or Ca(OH2) may be added

15 to the system either dry or as a Ca(OH)2 slurry. Desilication,

when using lime, is accomplished by heating the

aluminum-laden leachate to an elevated temperature

above about 100· C., preferably from about 180· C. to

abeut 200· C. To this heated leachate is added lime.

After a retention time sufficient to precipitate the aluminum

values present, preferably between about 5 and

about 30 minutes, the precipitated desilication product

(DSP) is separated from the leachate. While some desilication

will occur at temperatures below 180· C., typically

more vigorous temperature conditions are necessary

in a potassium system than in a sodium system.

Thus, the preferred temperature range of about 180· C.

to •about 200· C. represents an optimizing of temperature

conditions. Also, with longer retention time more

desilication will typically be accomplished, but with

concomitant higher aluminum precipitation loss. The

amount of lime is typically not critical, except to the

extent that insufficient lime will result in insufficient

seed for desilication. Thus, the lime added to the aluminum-

laden leachate is preferably in the range of about

12 to about 20 grams per liter of leachate, based on a

silica content in the leachate of about 1.5 gil Si02. The

lime is preferably added as an aqueous calcium hydroxide

slurry to maximize reactivity.

Following desilication, a high grade aluminum trihydroxide

is precipitated from the primary leachate and at

least some of the remaining liquor is advantageously

returned to the primary leach. The aluminum trihydroxide

may then be calcined to form the alumina product.

Potassium and aluminum values present in the ore

which have been rendered water soluble by the primary

leach remain in the residue due to the saturation of the

leach liquor with potassium sulfate and selection of

other primary leach conditions according to the present

invention. The solid residue, after separation from the

primary leachate is slurried with an aqueous solution,

preferably water which may contain at least some ofthe

spent liquor from potassium sulfate crystallization described

below, to solubilize potassium and sulfur values

present in the residue into the secondary leach liquor as

potassium sulfate. After separating the tailings from the

potassium sulfate-containing leachate, at least a portion

of this secondary leachate is treated by means known to

the art to produce potassium sulfate product, preferably

by crystallization. The remaining portion of the secondary

leachate may be treated to generate potassium hydroxide

in accordance with one or more of the caustic

generation processes described below. The potassium

hydroxide is then recycled to the primary alunite leach.

Bleed streams for controlling the level of impurities

including any sodium present from the ore and the

buildup of excessive potassium hydroxide in the secondary

leach circuit are provided according to the pro-

I Recovery of alumina and potassium sulfate from

alunite.

Alunite feed ores useful in the practice of the present

invention typically contain various amounts of silica 20

and other minerals. According to the present invention

there is provided a process for the recovery of aluminum,

potassium and sulfur as alumina and potassium

sulfate from alunite. The process involves leaching raw, 25

Le. thermally untreated, alunite ore with a potassium

hydroxide solution saturated with potassium sulfate

whereby aluminum is solubilized into the potassium

hydroxide and potassium sulfate-containing first leachate,

leaving the potassium and sulfur values of the ore in 30

the residue, but rendering them water soluble. Although

thermal treatment or roasting is unnecessary for

practice of the invention, as will be understood, roasted

ore could similarly be treated. The primary leach of the

present invention utilizes a strong potassium hydroxide 35

leach liquor which is saturated with K2S04, simultaneously

with or preferably prior to contacting with the

ore for primary leaching, and to which no.sodium has

been added as NaOH or otherwise. To the extent any

sodium is present, it will be that which was initially 40

present in the ore. Sodium build-up is controlled

through impurities bleed streams described herein

below and in all instances will be maintained below a

level of 70%.

The alunite ore is typically ground and/or crushed to 45

make it more amenable to leaching. The primary leach

must be conducted under temperature conditions that

are not too hot, and for a time that is not too long,

otherwise severe silica problems will result; Le., the

aluminum values may precipitate out of the leach solu- 50

tion as aluminosilicates and/or undesirably high levels

of silica will be extracted into the liquor. The primary

leach will typically be at a temperature above 60· C., at

least above 70· C., preferably, the primary leach temperature

is between about 70· C. and about 150· C., and 55

more preferably between about 90· C. and about 100·

C., and the retention is between about I minute and

about 120 minutes, and more preferably between about

20 minutes and about 40 minutes. Potassium hydroxide

concentration of the starting leach solution is between 60

about 180 gil and about 280 gil, preferably between

about 225 gil and about 245 ; gil and no sodium is

present. By selection of proper primary leach conditions

the following results are obtained: (a) maximizing

of the solubilization of the ore's aluminum values into 65

the leachate; (b) rendering most of the ore's potassium

values water soluble; (c) minimizing solubilizing of the

ore's potassium values into the leach solution; and (d)

regeneration scheme and thereby result in novel and

advantageous methods for metal value recovery.

BRIEF DESCRIPTION OF THE DRAWING

FIG. 1 is a general diagrammatic representation of

one embodiment of the present invention.

FIG. 2 is a phase stability diagram relating to the

conversion of potassium sulfate to potassium carbonate

by pyrohydrolysis.

FIG. 3 is a series of phase stability diagrams exemplifying

the effect of temperature.

FIG. 4 is a plot of aluminum extraction versus time at

60· C., 80· C. and 100· C.

DETAILED DESCRITPION OF THE

PREFERRED EMBODIMENTS

M2S04+Ca(OHh+ 2CO_2MCOOH+CaS04. (I)

2MCOOH+02-M2C03+C02+H20. (2)

The resulting alkali metal carbonates are sent to causticization

with lime in an aqueous solution where the corresponding

alkali metal hydroxides and solid calcium

carbonate are formed. The hydroxide solution is separated

from the precipitated calcium carbonate, with the

calcium carbonate precipitate advantageously sent to

lime regeneration and the hydroxide solution advantageously

recycled to the primary ore leaching step, after

evaporation, if necessary to achieve the desired concentration.

In a novel and advantageous process, alkali metal

sulfates may be converted to alkali metal hydroxides by

a direct hydrometallurgical route. This direct hydrometallurgical

route may be utilized in any process

wherein it is desirable to convert alkali metal sulfates to

the corresponding hydroxides.

In the process according to this invention, the alkali

metal sulfate-containing leachate is contacted with lime

in an aqueous solution at elevated temperature and pressure

in the presence of carbon monoxide gas and for a

time sufficient to produce the alkali metal formate. The

formate thus formed is oxidized to alkali metal carbonate

with an oxidizing agent, preferably an oxidizing gas

such as 02 or air. Upon conversion, the carbonate is

immediately reacted with lime to produce the alkali

metal hydroxide. Advantageously and preferably, the

lime is already present in the system as excess hydrated

lime over that amount consumed during the formation

of the formates. In another embodiment, the excess lime

may be added to the system during both or either the

carbon monoxide and/or oxidizing phases of the process.

The reaction steps may be represented as follows:

(M=alkali metal)

Temperatures suitable for the conversion reaction will

typically be from about 180' C. to about 260' C.; preferably

from about 210' C. to about 230' C. Pressures will

typically be in the range of from about 400 psi to about

700 psi.

After the formation of the potassium formate and

gypsum, the gypsum, CaS04, is separated from the

liquor, and is typically acceptable for disposal without

further treatment.

The alkali metal formates are crystallized from the

liquor remaining after the gypsum separation, by any

conventional method, preferably evaporation. The formate

crystals are then calcined at a temperature of from

about 350' C. to about 450' C., preferably about 400' C.

and for a time sufficient to produce the corresponding

alkali metal carbonates. The reaction is as follows:

The following described caustic generation processes

involving intermediate alkali metal formate and carbono

ate production are applicable to any process where it is

desirable to leach an ore with alkali metal hydroxides

for the recovery of metal values therefrom, which ore

contains alkali metal and sulfur values.

In one embodiment of the formate route for alkali

metal hydroxide regeneration, alkali metal sulfates are

contacted with hydrated lime, Ca(OHh at elevated

temperatures and pressures in the presence of carbon

monoxide gas for a time sufficient to convert the sulfates

to the corresponding formates, according to the

following reaction:

4,618,480

cesses of this invention. Separate bleed streams may be

taken from alternative points along both the primary

leach circuit and the secondary leach circuit. The bleed

streams can be treated by a variety of known methods

to remove impurities. In a preferred method, the bleed 5

stream from the secondary leach circuit is returned to

the primary leach circuit without treatment. In another

preferred embodiment, the bleed stream from the primary

circuit is reacted in a carbonation reactor to precipitate

aluminum and potassium hydroxide. Excessive 10

potassium hydroxide in the primary leach causes a deterioration

of the conditions for desilication from the

alumina-containing primary leachate. Furthermore,

excessive potassium hydroxide in the secondary leach

solution can cause a decrease in the solubility of sulfate, 15

hence, less productivity in the potassium sulfate crystallization

circuit.

2. Hydrogen sulfide-lime caustic generation.

The following described caustic generation process is

applicable not only in the above-described process for 20

the recovery of alumina and potassium sulfate from

alunite, but to any process where it is desirable to leach

an ore with one or more alkali metal hydroxides for the

recovery of metal values therefrom, which ore contains

alkali metal and sulfur values. 25

The conversion of alkali metal sulfates to alkali metal

carbonates is conducted generally according to the

process taught in U.S. Pat. No.3, 134,639. Hot saturated

alkali metal sulfate solution is reacted with lime slurry

and hydrogen sulfide gas, e.g. in a spray tower, to form 30

gypsum as a precipitate and soluble alkali metal sulfides.

These alkali metal sulfides include alkali metal hydrosulfides.

A liquid/solid separation is performed and the

gypsum tailings are sent to disposal. The alkali metal

sulfide-containing liquor is next treated with carbon 35

dioxide, preferably in an absorption tower, to form

alkali metal bicarbonates and carbonates and hydrogen

sulfide. The hydrogen sulfide is preferably recycled to

treat additional alkali metal sulfates. The bicarbonate/carbonate

solution is then contacted with water at high 40

temperature, preferably by steam stripping, whereupon

carbon dioxide is removed. This carbon dioxide is preferably

recycled to treat the alkali metal sulfides as described

above. Approximately half the required carbon

dioxide is generated by this method. During the treat- 45

ment of the bicarbonate/carbonate solution to release

carbon dioxide, the bicarbonates are converted to carbonates.

The alkali metal carbonate solution is then

causticized to generate alkali metal hydroxide therefrom,

preferably by reacting with lime at about 85' C. to 50

about 95' C. The resultant solution desirably contains

about 12% alkali metal hydroxide. During treatment

with lime, calcium carbonate is precipitated. A liquid/solid

separation is performed and the calcium carbonate

is advantageously heat treated in the presence of water 55

to generate lime for recycle to the first step of the conversion

process and/or the causticization process, and

carbon dioxide for recycle to the treatment ofthe alkalimetal-

sulfide-containing liquor. The alkali metal hydroxide

solution is advantageously recycled to the pri- 60

mary ore leaching step, after evaporation, if necessary,

to achieve the desired concentration. It has been discovered

that utilization of this regeneration process in conjunction

with the novel KOH alunite leaching methods

of the present invention advantageously results in a 65

novel overall process wherein inexpensive CaO is the

sole consumptive reagent.

3. Formate caustic generation.

(8)

(7)

(3)

(6)

(4)

4,618,480

M2S04+2CO+Ca(OH12-.2MCOOH+CaS04

M2S04+0 2+2CO+2Ca(OH)

2-.2MOH+CaS04+CaC03+C02 +mO.

K2S04+CH4-.K2C03+H2S+H20. (9) 65

The carbonates are then causticized with lime as described

above to form the corresponding hydroxides.

The formed carbon dioxide may be recycled to the

M2C03+Ca(OHh-.2MOH+CaC03. (5) 5

Thus, the net reaction may be exemplified as follows:

reactor if desired. The above reaction may be conducted

in one stage in a furnace or fluid bed reactor, as

described in the examples hereof.

When coal is used as the reducing agent, a stoichiometric

ratio of carbon to alkali metal sulfate of between

about 1 and about 3, preferably between about 1.1 and

about 1.3 is desirable.

When reducing gases are used, these may be selected

from the group consisting ofhydrogen and carbon mon10

oxide, and hydrocarbons which are in the gas phase at

The reaction for converting alkali metal sulfates to the reaction temperature. Of these, the low molecular

alkali metal hydroxides may be conducted within the weight aliphatic alkanes are probably of most interest. It

temperature range of about 180· C. to about 280· C.; is essential that the gases selected provide both carbon,

preferably about 200· C. to about 240· C. The pressure for the formation of carbonates, and hydrogen for the

for the reaction is typically from about 200 psi to about 15 formation of H2S. An example of the hydrocarbon gas

1500 psi; preferably from about 300 psi to about 600 psi. reaction is shown in reaction (9) above. As will be un-

The carbon monoxide is added to the system, during the derstood by those skilled in the art, equilibrium condiformate

production, preferably by sparging. The reac- tions among the gaseous reactants will be established

tion time for formate production under the carbon monoxide

atmosphere is typically from about 2 to about 60 20 during the reaction, dependent upon the temperature

minutes. The oxidizing agent is preferably an oxidizing and pressure of the reaction, and these should be such

gas. The reaction time under the oxidizing atmosphere that the stable solid phase in equilibrium with this gas

is typically from about 2 to about 60 minutes. phase is K2C03 an example of which is shown in FIG.

In preferred embodiments, the formate caustic regen- 2. The boundaries of the K2C03 field change with temeration

methods are used in conjunction with the novel 25 perature, becoming more limited as temperature is in-

KOH alunite leaching ofthe present to advantageously creased as shown in FIG. 3.

generate the potassium hydroxide from potassium sul- Temperature is a criii~al parameter and should be

fate for recycle to a primary alunite leach, as described between about 600· C. and about 1000· C., preferably

above. The potassium sulfate contained in the second- between about 750· C. and about 850, namely high

ary leachate is reacted with hydrated lime at elevated 30 enough to allow reasonable kinetics without entering

temperatures and pressures under a carbon monoxide the region of fusion and appreciable vapor pressure of

atmosphere to produce potassium formate. Next, potas- the solid reactants and products.

sium carbonate is. formed, but in the presence of an Reaction time for reactions conducted in a furnace

·oxidizing agent. The thus-formed potassium carbonate should be between about 15 minutes and about 5 hours,

is converted to potassium hydroxide by the contact of 35 preferably between about 30 minutes and about 1 hour.

the potassium carbonate with hydrated lime. The potas- When the reaction is conducted in a fluid bed reactor,

sium hydroxide-containing solution is separated from reaction time should be between about 15 minutes and

the precipitated calcium sulfate and calcium carbonate, about 5 hours, preferably between about 15 minutes and

typically by filtration. The potassium hydroxide solu- about 1 hour.

tion is recycled for use in the ore leaching. In the alunite 40 In a preferred embodiment, coal is burned in the

processes which include regeneration of CaC03, cal- presence of air and steam to provide both the heat for

cium sulfate in the secondary leachate is advanta- the reactions and the reducing gases used as feed to the

geously separated, e.g. by solidlliquid separation prior reactor.

to either K2C03 or CaC03 formation. By the process of this invention, up to 98.6% conver-

4. Pyrohydrolysis caustic regeneration. 45 sion of alkali metal sulfates to the corresponding car-

The novel pyrohydrolysis conversion of alkali metal bonates can be achieved.

sulfates to hydroxides is useful not only in connection 5. Barium oxide alkali metal hydroxide regeneration.

with the above-described alunite process, but in connec- The following described process is suitable not only

tion with any process in which it is desired to convert for use with the above-described potassium hydroxide

alkali metal sulfates to alkali metal carbonates and/or 50 alunite leach of the present invention, but also may be

hydroxides, including processes for leaching of ore used in connection with alkali metal hydroxide leaches

materials with alkali metal hydroxides, in which the ore of any ores wherein it is desired to convert alkali metal

materials contain alkali metal and sulfur values which sulfates to the corresponding hydroxides for recycle to

may be recovered during processing of the ore as alkali the leach, preferably wherein the alkali metal and sulfur

metal sulfates and converted to the corresponding hy- 55 values required are found in the ore itself. The process

droxides for recycle to the ore leach. may be used, for instance, in a combined sodium hy-

In the process of this invention an alkali metal sulfate droxide/potassium hydroxide leach of alunite or other

is reacted with coal and/or a reducing gas to produce ores.

the corresponding carbonate according to the following In the process of this embodiment, alkali metal sulfate

reactions: 60 is reacted with barium oxide and water (the barium

oxide may be pre-hydrated to form the hydroxide or

added to the solution as unhydrated barium oxide) to

directly form the alkali metal hydroxide and insoluble

barium sulfate. After a liquid/solid separation step, the

barium 'sulfate is then reacted with a carbonaceous fuel

at an elevated temperature, preferably around 1200· C.

to form solid barium sulfide and carbon dioxide,. which

products react together with water to form a barium

4. Secondary Leach

In the preferred embodiment, the residue 13 is slurried

and/or leached with water to solubilize the potassium

and sulfur values contained in the residue. Advantageously,

spent liquor 18 returning from the K2S04

precipitation 5, described below, may comprise part of

the slurried solution. The secondary leach 4 is typically

a two-stage countercurrent leach wherein the potassium

and sulfate values solubilize into the liquor secondary

leachate 14. The secondary leachate 14 and remaining

solid ore residue tails 15 are separated, typically by

thickening and filtration. The residue tails 15 may be

further washed, e.g. with water, to recover additional

soluble potassium and/or sulfur values. The washed

non-toxic residue tails 15 from this leach may be sent to

a tailings pond for disposal. In a preferred embodiment,

the secondary leachate 14 is divided into at least two

3. AIz03 Recovery

The primary leachate 10, i.e. the aluminum-laden

liquor from the primary leach, contains dissolved silicates.

It is desirable to obtain a pure grade alumina from

the leachate by direct crystallization. However, the

presence of dissolved silicates in the leachate would

result in unacceptable silica contamination of the crystals

and thus the leachate must be be desilicated prior to

crystallization. The leachate is typically supersaturated

and as such requires seeding in order to desilicate. According

to the present invention effective desilication of

the leachate is accomplished by contacting or seeding

the leachate with CaO to precipitate out the silica as

insoluble silicates.

Desilication is accomplished by heating the aluminum-

laden leachate above about 100' c., preferably to a

temperature offrom about 180' to about 200' C. To this

heated leachate is added CaO, typically as a Ca(OHh

slurry. The Ca(OHh concentration in the a1uminumladen

leachate is from about 12 to about 20 grams of

Ca(OHh per liter of leachate. The retention time is from

about 5 to about 30 minutes. After the retention time,

the precipitated desilication product (DSP) 11 is separated

from the leachate and sent to disposal.

The silicate precipitant is separated from the leachate

liquor, typically by filtration. The resulting liquor may

then be processed according to methods known in the

art. Typically, the liquor is sent to aluminum trihydrate

precipitators followed by slurry classification and washing.

The course hydrate formed during trihydrate precipitation

is thickened, filtered, washed, and calcined to

form the final AIz03 product. The fine hydrate is returned

to the precipitators for seed AI(OHh. The KOH

liquor 12 remaining after the aluminum trihydrate precipitation

is advantageously recycled to the primary

leach 2.

time and temperature can vary inversely. At higher

temperatures leaching is for shorter periods of time, e.g.

130' C. for 3 minutes. However as will be known and

understood by those skilled in the art, loss of aluminum

as precipitated aluminosilicates can be balanced by

other overall factors and thus technical operability of

the leach system can be achieved over the entire ranges

of temperatures and times provided. The concentration

of KOH in the leachate 10 after the primary leach 2 is

10 from about 140 to 220 grams KOHlliter, preferably

from about 160 to about 200 grams KOHlliter and

typically about 180 grams KOHlliter.

(10)

4,618,480

K2AI6(504).(OHh2 + 6KOH~+ 6Al(OHh +4K2504.

It has been discovered that the aluminum values can

be selectively taken into solution leaving the potassium

values from the ore in the leach residue by utilizing a

KOH leach solution which is initially saturated with

potassium sulfate prior to the leaching 2. Additionally, 40

it has been discovered that by contacting the ore with a

primary leach comprising K2S04-saturated KOH, the

potassium values in the residue, although not solubilized

into the leachate, will nevertheless be rendered water

soluble and thus easily recoverable in a subsequent wash 45

or leach. Moreover, by selection of proper leach conditions

of time and temperature above 60' C., minimal loss

of aluminum values due to precipitation of aluminosilicates

can be achieved despite the substantial amount of

silica in the prepared ore. 50

The concentration of K2S04 in the leach solution is

sufficient to saturate the solution under the particular

conditions, but will typically be from about 20 to about

30 grams K2S04lliter. The KOH leach liquor may

advantageously comprise at least in part the spent liquor 55

18 returning from the K2S04 recovery 5. Approximately

90% of the aluminum values are solubilized in

the primary leach 2. Also, 95% of the sulfur and potassium

values of the alunite ore are rendered water soluble,

but remain in the residue 13. The liquor 10 contain- 60

ing the dissolved aluminum is sent to alumina recovery

3 and the residue 13 is sent to a secondary leach 4. The

primary leach 2 is conducted at a temperature above 60'

C., typically from about 80' C. to about 150' C., preferably

at about 95' C., with a retention time of from about 65

2 minutes to about 2 hours, preferably about 30 minutes.

To minimize loss of aluminum values due to precipitation

of insoluble aluminosilicates, the conditions of

carbonate precipitate and hydrogen sulfide offgas. The

barium carbonate precipitate is then reacted with coal

or coke at an elevated temperature, preferably about

1100' C. to regenerate barium oxide for recycle to the

process and carbon monoxide which may also be recy- 5

c1ed to the process.

Referring to FIG. 1, which is a general diagrammatic

flow sheet of the preferred embodiment of the invention,

the alunite ore is processed in the following steps.

1. Ore Preparation

The alunite feed ore is advantageously physically

reduced in size by crushing and/or grinding in ore preparation

1, e.g. crushed to approximately minus a-inch

material and ground to approximately 20-mesh. In a

preferred embodiment, a pre-leach solution of KOH is 15

added to the final grinding of the crushed/ground ore

product during which an approximately minus 20-mesh

Tyler product is achieved. This KOH-preleach solution

may advantageously be th€ spent leach liquor from the

primary leach 2, described below. The pre-leached 20

slurry is then sent to the primary leach 2.

2. Primary Leach

The ore slurry from the preparation 1, undergoes a

primary leach 2, typically a single stage leach, with a

strong KOH leach solution, saturated with K2S04. For 25

purposes of this invention a "strong" KOH leach is one

containing a concentration of KOH remaining in the

spent primary leach from about 160 to about 240 grams

KOHlliter of solution, preferably from about 180 to

about 200 grams KOHlliter, and most preferably about 30

180 grams KOHlliter. The key reaction taking place

during the primary leach is:

Element %

AI 9.74

S 7.53

K 4.17

Na 0.230

Ca 0.094

Mg 0.017

Fe 0.742

Si 23.0

8. Lime Regeneration

In the .preferred embodiment, regeneration of.lime,

CaO 8, from precipitated CaC0324 is accomplished by

5 calcining the CaC03 limestone formed during causticization

7, typically at a temperature of from about 800°

to about 1200° C. and for a time sufficient to convert

CaC03 to CaO. The CaO 25 is then recycled for use in

the causticization 7.

4,618,480

Example 1

A series of experiments was performed on a sample of

Utah alunite ore. The sample is believed to be represen55

tative of the high grade core ofthe NG ore deposit near

Cedar City, Utah. Table 1 shows the chemical analysis

7. Causticization - KOH Regeneration of this sample. The sample was stage crushed to minus

In one preferred embodiment, causticization 7 of 20 mesh and dried for 24 hours at 110° C. in preparation

aqueous K2C03 by reaction with lime (CaO) is effected for conducting the experiments.

to produce KOH. The causticization is performed on 60 TABLE 1

the aqueous K2C03 utilizing countercurrent methods, --------------------

preferably at a temperature of from about 85° to about

95° C. The products of causticization are a solution of

approximately 12%-15% KOH 23 and precipitated

limestone, CaC03 24. The CaC03 precipitate 24 is sent 65

to lime regeneration 8, described below, and the KOH

solution 23 is advantageously recycled to the primary

leach 2.

streams 16 and 17, with one stream 16 being sent to

K2S04 crystallization 5, and the other portion 17 being

sent to K2C03 generation 6, both of these steps being

described below.

9. Impurities Control

A bleed stream 26 for controlling impurities of the

primary leach 2 is taken from the KOH liquor 12 remaining

after AI203 recovery 3 and before recycling

this liquor 12 back to the primary leach. An alternative,

and/or additional, point at which a bleed stream may be

taken in the primary leach circuit is 27, in which a bleed

stream from the alumina-containing primaryleachate 10

is sent to impurities control 9. The treated bleed stream

28 may then be returned to the primary leach 2.

A bleed stream from the secondary leach circuit may

be taken from either or both of two points in the secondary

circuit. A bleed stream 29 for controlling impurities,

primarily excessive KOH buildup, may be taken from

the K2S041iquor 18 returning to the secondary leach 4

after K2S04 recovery 5, with the bleed stream 29 returning

to the primary ieach 2. An alternative, and/or

6. K2C03 Generation additional, point at which a stream may be taken in the

In another preferred embodiment, a portion of the secondary leach circuit is 30, from a portion of the

secondary leachate solution 17 containing dissolved 30 K2S041eachate 14. This bleed stream is also sent to the

potassium and sulfate is processed to generate K2C03 primary leach 2.

and ultimately to regenerate KOH. The secondary FIG. 2 shows a stability diagram for the K-O-H-C-S

leachate 17, typically containing about 12% to about system of the pyrohydrolysis KOH regeneration de-

18% K2S04, is reacted with a lime slurry (Ca(OHh) 20 35 scribed herein at 15% C02 and Pt= 1 atm. Referring to

and H2S gas in a gaslliquid contacting device capable of FIG. 2, the log of the pressure of H2S divided by the

handling solids, e.g. a spray tower, to form solid gyp_ pressure of H2 is plotted on the Y axis. The log of the

sum, CaS04.2H20, and soluble K2S and KHS. The H2S pressure ofC02 divided by the pressure ofCO is plotted

gas concentration is not critical. Contact with K2S04 on the X axis. For a given temperature, such a stability

results in the formation of gypsum which is separated 40 diagram can be constructed. In order to produce

from the spent liquor typically by countercurrent de- K2C03, the equilibrium gas composition in a system

must fall within the K2C03 stability field given by a

cantation (CCD) followed by vacuum filtration. The diagram such as FIG. 2. The selected temperature is

g)1psum tailings 24 can be sent to a disposal pond. The desired to be below the fusion point of the K2S04 and

K2S and KHS liquor is sent to an absorption tower K2C03, and above about 700° C.

where C02 is absorbed forming KHC03 and H2S. The 45 As will be understood by those skilled in the art,

concentration of C02 typically required for this step is modifications of the above process may be made withabout.

20%. The H2S is ~ecycled to the spray tower out departing from the scope of the invention. The

deSCribed above for forming gypsum and soluble K2S. following examples are provided for illustration and not

The KHC03 solution is steam stripped to release one- by way of limitation.

half of the C02required for K2S carbonation, the bal- 50

ance of C02 being supplied from the lime regeneration

8, described below. Aqueous K2C0321 is formed when

the KHC03 is steam stripped. The K2C03 21 is then

sent to the causticization 7, described below.

5. K2S04 Recovery

In the preferred embodiment, a portion of the secondary

leachate 16 from the secondary leach containing

dissolved potassium and sulfate is treated for recovery

of solid K2S04 by crystallization 5. The crystallization 10

may be by any means known in the art, typically by

utilizing a vacuum cooled crystallizer, operating at a

crystallization temperature of about 40° C. As will be

known and understood by those skilled in the art, the 15

leachate from the secondary leach may have varying

amounts of K2S04, e.g. from about 12% to about 18%

K2S04. After crystallization, the crystallized K2S04 19

is separated from the spent liquor by means known in

the art. The spent liquor 18 typically containing about 20

8% to about 12%, more typically about 10% residual

K2S04 is advantageously returned to the secondary

leach circuit 4. The crystallized K2S04 19 may be further

processed according to means known in the art,

such as centrifuging, compacting, and drying for com- 25

mercial use.

Example 2

Leach tests were conducted to examine the effects of

time, temperature, and KOH concentration on aluminum,

potassium and sulfur extraction. Most of the leach

tests conducted report aluminum, potassium, and sulfur

extraction results. The primary KOH leach is intended

to solubilize only aluminum, not potassium and sulfur.

During the KOH leach, however, the potassium and

TABLE I-continued

Element

Total organic carbon (TOC)

0.002

1.81

4,618,480

sulfur in the residue are rendered water soluble (as free

potassium sulfate). The leaching technique used in the

tests comprised filtration of the slurry after agitation at

specified time and temperature, followed by at least

5 four water washes. The water soluble K2S04 is thus

solubilized and usually collected with the aluminum

bearing primary filtrate. Hence total aluminum, potassium,

and sulfur extractions are reported in a single

leach. These are considered representative of total ex-

10 tractions attainable in the two-step leach process.

Table 2 summarizes conditions and results for the

leach tests performed. Aluminum extraction results

versus time at 600 C., 800 C. and 1000 C. are plotted in

FIG. 4.

TABLE 2

Leach Test Conditions and Results

No. Feed

Conditions

Initial

KOH Cone X Stoich

Temp

"C.

Time

min Al

Extraction (gil)

S K Na

Available

Alumina

No reaction

75.9 74.5

A Raw ore

B Dehydrated

ore

C Raw ore

F Raw ore

I Raw ore

L Raw ore

(pH 10.0)

10% 1.4

18% 2.8

10% 1.4

18% 2.8

130

150

120

11.4

0.0

34.7

27.3

82.1

83.3

23.1

30.3

30.2

26.5

25.6

13.3

13.2

15.3

51.9

41.3

90.3

94.5

40.5

44.5

45.4

42.0

47.2

60.4

16.5

18.1

43.5

42.0

89.4

93.8

39.7

44.4

45.4

43.4

47.0

57.5

25.6

28.5

55.0

93.2

93.8

89.2

91.3

60.2

91.3

91.4

89.8

90.4

62.5

Leach

No.

KOH

Cone. gil

(X Stoich)

Temp

"C.

Time

min Sample Al

Assay. % or gil

S K Na Si

% Extraction

(Accountability)

Al S K

35.6 32.4

178 (4.1)

174 (4.1)

171 (4.1)

167 (4.1)

231 (4.4)

226 (4.4)

221 (4.4)

220 (4.4)

209 (4.2)

100

15 Mother liquor

Head

Residue

Filtrate & wash

30 Mother liquor

Head

Residue

Filtrate & wash

60 Mother liquor

Head

Residue

Filtrate & wash

120 Mother liquor

Head

Residue

Filtrate & wash

15 Mother liquor

Head

Residue

Filtrate & wash

30 Mother liquor

Head

Residue

Filtrate & wash

60 Mother liquor

Head

Residue

Filtrate & wash

120 Mother liquor

Head

Residue

Filtrate & wash

15 Mother liquor

Head

Residue

28.8

10.0

2.24

12.6

28.8

10.0

1.41

10.9

28.8

10.0

1.13

12.6

28.8

10.0

1.04

11.1

28.6

9.81

6.82

28.6

9.81

7.19

28.6

9.81

8.83

28.6

9.81

8.83

28.6

9.81

7.82

0.715 0.187

7.38 3.98 0.225

1.13 0.617 0.059

4.57 0.613

0.715 0.187

7.38 3.98 0.225

0.463 0.308 0.047

4.72 0.165

0.715 0.187

7.38 3.98 0.225

0.317 0.301 0.046

4.87 0.172

0.715 0.187

7.38 3.98 0.225

0.253 0.257 0.045

4.35 0.154

0.195

7.25 3.87 0.227

5.91 3.30 0.212

0.031

23.0

0.05

0.031

0.056

0.031

0.100

87.5

(100.2)

96.4

(85.8)

97.1

(96.3)

97.4

(95.6)

6.1 1

91.8

(100.6)

96.8

(99.4)

98.1

(99.8)

98.3

(100.3)

91.4

95.9

96.0

96.6

29.0

4,618,480

TABLE 2-continued

Leach Test Conditions and Results

Filtrate & wash 7.85 1.76 (98.0) (107.6)

10 214 (4.2) 80 30 Mother liquor 28.6 0.195

Head 9.81 7.25 3.87 0.227

Residue 6.74 5.04 2.64 0.148 47.8 47.0 48.0

Filtrate & wash 8.68 2.05 (97.1) (98.7)

II 197 (4.2) 80 60 Mother liquor 28.6 0.195

Head 9.81 7.25 3.87 0.227

Residue 4.42 2.88 1.58 0.100 71.4 74.9 74.1

Filtrate & wash 9.85 3.76 (94.3) (107.9)

12 179 (4.2) 80 60 Mother liquor 28.6 0.195

Head 9.81 7.25 3.87 0.227

Residue 2.14 0.909 0.606 0.059 88.1 93.0 91.4

Filtrate & wash 12.1 4.35 (102.8) (101.6)

13 197 (5.4) 100 30 Mother liquor 28.6 0.195

Head 9.81 7.25 3.87 0.227

Residue 1.41 0.311 0.410 0.049 92.5 98.0 94·1

Filtrate & wash 11.6 3.52 0.143 0.054 (106.0) (98.8)

14 176 (3.5) 100 30 Mother liquor 28.6 0.0 0.194 0.029

Head 9.81 7.25 3.87 0.227

Residue 1.95 0.692 0.406 0.056 89.2 94.9 94.3

Filtrate & wash 12.0 5.32 0.155 0.061 (92.3) (100.8)

15 185 (3.8) 100 30 Mother liquor 28.6 0.0 0.195 0.029

Head 9.81 7.25 3.87 0.227

Residue 1.67 0.528 0.316 0.053 90.8 96.2 95.6

Filtrate & wash 13.61 4.86 0.172 0.060 (106.2) (98.0)

lCalculated from the filtrate assay.

2Samples in leach nos. 5-8 were analyzed only for aluminum since it was apparent from weight loss and A/K titration results that poor extractions

had been achieved.

Example 3 TABLE 4

Leach tests were conducted to determine the effect of 30 Assay, % or g/l % Extraction

KOH concentration on Si02 attack and on aluminum Sample Al S K Na Si Al S K

extraction. The results are shown in Table 3. Mother 26.0 7.39 0.190 0.06

TABLE 3

Leach Test Conditions and Results

KOH % Extraction

Leach Cone. gil Temp Time Assay, % or g(] ('Balance, %)

No. (X Stoich) 'c. min Sample Al S K Na Si AI S K

160 (3.4) 100 30 Syn. spent liquor 21.2 0.178 0.08

Head 10.3 7.46 4.13 0.281

Residue 2.11 0.84 0.47 0.069 88.7 94.7 93.7

Filtrate & wash 12.8 6.46 0.212 0.45 (103.3)

2 200 (3.4) 100 30 Syn. spent liquor 27.6 0.245 0.13

Head 10.3 7.46 4.13 0.178 0.08

Residue 1.86 1.27 1.76 0.050 88.9 86.1 76.1

Filtrate & wash 16.7 7.78 0.252 0.74 (104.4)

220 (3.4) 100 30 Syn. spent liquor 31.2 0.245 0.12

Head 10.3 7.46 4.13 0.281

Residue 0.680 0.66 0.47 0.51 90.5 95.3 93.9

Filtrate & wash 17.7 8.70 0.288 0.91 (98.6)

4 180 (3.4) 100 30 Syn. spent liquor 25.1 0.202 0.14

Head 10.1 7.46 4.13 0.281

Residue 1.98 0.79 0.46 0.053 84.4 94.2 93.9

Filtrate & wash 14.4 6.92 0.231 0.56 (98.6)

'Mass balance based upon AI20)/KOH ratio which for the leaches gave a more accurate number than when based upon the Al analysis.

liquor

Head 10.1 3.97 0.241

55 Residue 1.96 0.93 0.89 0.63 89.3 93.2 87.8

Filtrate 9.64 3.5 - 0.141 0.09

& (Accountability, %) (97.6) (94.6)

wash

Example 4

A leach test was conducted to determine the result of 60

K2C03 contamination of the leach liquor. A standard

leach was conducted with the leach KOH concentration

of 197 gil (KOH stoichiometric multiple 4.3) and

20% K2C03; temperature 100· C.; and leach time 30

minutes. The test showed that aluminum extraction is 65

decreased and that Si02 extraction is increased by the

addition of 20% K2C03 to the leach liquor. The results

are shown in Table 4.

Example 5

A leach test was conducted to determine the effect of

long holding time on extraction of values and Si02

levels. A standard leach was conducted with the leach

KOH concentration 233 gil (KOH stoichiometric multiple

4.3); temperature at 100· C.; and leach time 30

minutes. The resulting slurry was maintained at 80· C.

for 24 hours. The purpose of this approach was to deter·

Conditions: 100° C. for 30 minutes. final KOH. approximately 190 gil,

••All figures rounded off.

mine iflong holding time in thickeners might be a problem

due to desilication reactions. Aluminum extraction

was not affected, but Si021evel in the liquor increased.

The results are shown in Table 5.

TABLE 8-continued

TABLE 5

Assay. %or gil %Extraction

4,618,480

A/K Final Filtrate

Target Ratio

0.7

Si02 gil

in filtrate

0.59

Sample Al S K Na Si Al S K

26.0 7.39 0.190 0.06 EXAMPLE 7

10 10.1 3.97 0.241 Aluminum hydroxide precipitate was spectrochemi-

1.34 0.207 0.365 0.049 93.1 98.6 95.2 cally analyzed and the results compared with spectro-

10.9 4.15 0.170 0.16 chemical analyses of the ore for the same elements.

(Accountability. %) (92.9) (92.9) Results are set forth in Table 9.

Mother

liquor

Head

Residue

Filtrate

wash

...-,;-------------------- 15 TABLE 9

0.1

0.0003

0,0)

Major

0.03

0.01

Precipitate

Analysis

0.1

0.003

0.005

Ore

Major2

0.1

4.0

0.3

0.002

0.01

0.001

0.1

EXAMPLE 8

Semiquantitative Emission Spectrochemical

Analysis of Ore and AI(OHh Precipitate

Element

Silicon

Aluminum

Iron

Calcium

Magnesium

Sodium

Titanium

Manganese

Chromium

Copper

Nickel

Lead

Zinc

Molybdenum

Vanadium

Strontium

Barium

lAnalysis expressed as a weight percentage· estimate only.

2Major represents a concentration above 10%.

0.52

0.55

Si02 gil

in filtrate

TABLE 8

0.5

0.6

AIK Final Filtrate

Target Ratio

A/K Final Filtrate

Target Ratio Al% K% S%

0.5 92 95 98

0.6 91 95 96

0.7 89 95 95 50

'Conditions: 100' C. for 30 minutes. final KOH. approximately 190 gil.

••All figures rounded off.

A series of tests was performed to determine the

impurity buildup on recycling use of the leach liquors.

The tests were conducted in a manner to simulate the

parameters outlined in the detailed description portion

of this specification. This testing did not, however,

include regeneration of KOH from K2S04. Thirteen

45 cycles utilizing the leach liquors were conducted and

---....,..----------------...... evaluated.

The test conditions were as follows: In Cycle 1, the

alunite ore was leached for 30 minutes at 100· C., filtered,

and washed three times with H20. The residue

was repulped with 400cc H20 at 90· C., filtered, and

washed with 40· C. H20. Reagent Al(OH)3 was added

as a seed to the pregnant liquor. This was mixed for 22

hours at 60· C., filtered, washed three times with 75cc

TABLE 7 H20 followed by a separate H20 wash. In Cycles 2-12

--------------------- 55 the ore was leached with spent liquor for 30 minutes at

AIK Final Filtrate 100· C.; then cooled to 60· C.; filtered and washed with

Actual Ratio AI %

60 cc H20. The leach residue was repulped with barren

0.5 92 K2S04 liquor (excess K2S04 filtered out) at 90· C.,

0.6 ~~ filtered and washed with 40· C. H20. This filtrate was

_______0_.7 60 evaporated to approximately 200 cc, cooled to 40· C.,

'Conditions: 100' C. for 30 minutes. final KOH. approximately 190 gil. filtered and washed with H20. Al(OH)3 seed from the

**All figures rounded off.

previous cycle was added to the leach filtrate, mixed 22

hours at 60· C., filtered and washed with H20. In Cycle

13, 5 cc, 1 gil Polyox 301 was added to the leached ore

65 after it cooled to 60· C. The leach filtrate was boiled for

24 hours, filtered, and washed with H20. The leach

residue was repulped with barren K2S04 solution at 90·

C. This repulped residue was filtered, and the residue

Example 6

Tests were designed to study the effect of target A/K

ratio on the extraction of the AI, K, and S values. The 20

target A/K ratio is defined as the weight ratio of Ab03

to KOH which will result in the liquor if 100% of the

aluminum is extracted and KOH is consumed stoichiometrically

according to the reaction,

A!2(S04h+6KOH---+(2AI(OHh+3K2S04

The target A/K can be thought of as an ore charge ratio

for the leach. High final ratios (the final A/K ratio

achieved) are desirable for liquor productivity but must

be consistent with high extraction objectives.

Table 6 shows the test results in terms of AI, K, and

S extraction as a function of target A/K ratio. Table 7

shows Al extraction as a function of final ratio (AIK

ratio actually achieved). These results illustrate that

extraction of values is an inverse function with respect 35

to target A/K ratio, as well as final ratio, though the

effect is not dramatic in the range ofAIK ratios studied.

A slight increase in Si02 in the liquor results in increased

target A/K, as shown in Table 8. 40

The results indicate that the target A/K ratio is advantageously

chosen by economic analysis in addition

to technical aspects. A target ratio of 0.6 was used for

most of the standard tests in these Examples.

TABLE 6

4,618,480

again repulped with H20 at 90° C. for approximately 10

TABLE 11

minutes, then filtered and washed three times with H2O.

Results of the impurities build-up tests are summa- Analysis of Cycle 13 Spent Liquor

Stream 5

rized in Table 10. The spent liquor of Cycle 13, stream

Analysis 5, is given in Table 11. Analyses of the products are 5 Species ppm

summarized in Table 12. Ga 31

Definition of streams: B 135

Stream 1 is the leachate, containing the dissolved P 533

aluminum, following the primary KOH leach and filtra- Ca I

tion separation of the residue from the leachate. 10 Mg I

CI- II

Stream 2 is the liquor resulting from the separation of Ti I

the precipitated Al(OHh Cr 5.5

Stream 3 is the resulting liquor from the secondary Mn 0.5

leach of wash of the residue containing the potassium Ni 5.7

Fe 6.7

and sulfur values, after repulping and filtering the ex- 15 Pb 14.0

cess K2S04. CU 1.5

Stream 4 is the returning liquor following separation Zn 1.2

ofK2S04.

V 129

CZ04= (oxalate) 8.6

TABLE 10

Analysis of Liquors, gil

Cycle No. 2 4 5 6 7 8 9 10 II 12 13

Leach Al 51.0 48.9 57.1 50.8 48.1 49 49 50.0 42.5 46.3 50.0 43.7 48.6

liquor K 121 119 106 105 112 105 115 116 116 114 118 115 -

(Stream #1) Na 0.745 1.35 1.88 2.35 3.04 2.94 3.14 3.33 3.59 3.54 3.50 3.59 3.78

S 7.74 7.41 8.11 5.44 5.44 4.27 4.57 4.80 4.54 4.78 4.31 4.68 5.44

Si 0.27 0.45 0.86 I.Dq 0.82 1.09 1.0 1.0 1.05 1.12 1.09 1.04 0.97

Spent liquor Al 16.2 24.5 25.6 20.8 21.4 22.5 23.8 27.5 25.0 23.8 27.5 25.0 21.0

(Stream #2) K III 101 115 105 109 109 118 118 116 116 116 116

Na 0.763 1.38 2.03 2.45 2.70 2.84 3.04 3.33 3.59 3.50 3.35 3.52 3.83

S 5.30 5.04 4.34 5.34 4.87 5.04 5.43 4.74 4.54 4.78 4.04 4.54 5.41

Si 0.33 0.53 0.90 0.91 0.76 0.83 0.90 0.90 0.98 0.99 1.09 1.00 1.00

Repulp Al 1.49 3.26 4.43 5.48 6.41 6.37 6.62 7.24 6.32 6.32 7.30 7.15 6.18

PF + w K 53.6 84.6 77.2 75.0 77.3 77.7 78.6 74.6 72.3 77.2 76.0 77.0

(Stream #3) Na 0.034 0.109 0.161 0.253 0.346 0.391 0.497 0.542 0.545 0.580 0.600 0.638 0.108

'S 19.8 26.5 27.0 27.0 25.9 25.9 25.2 25.2 24.6 25.2 25.5 24.8 23.8

Si 0.03 0.01 0.05 0.11 0.01 0.06 0.01 0.05 0.05 0.08 0.12 0.12

Barren AI 1.49 3.96 6.76 9.09 9.09 8.09 10.3 11.1 9.75 9.25 11.O 11.3 7.78

KZS04 K 16.9 18.3 20.3 25.0 27.7 28.7 36.0 31.1 36.8 32.4 33.3 45.1

liquor Na 0.032 0.125 0.250 0.373 0.476 0.554 0.739 0.804 0.835 0.879 0.950 0.994 0.670

(Stream #4) S 5.67 3.77 0.208 2.04 2.89 2.76 3.06 1.56 3.54 1.39 1.19 1.79

Si 0.01 0.02 0.06 0.08 0.08 0.08 0.13 0.13 0.12 0.12 0.14 0.14

KOH 2.8 9.8 23.2 26.3 30.8 28.6 39.2 39.2 37.5 39.2 38.1 38.9 25.8

Total organic carbon 70

TABLE 12

Analysis of Products, Wt %

Cycle No. 2 3 4 5 6 7 8

Trihydrate SiOz 0.032 0.026 0.026 0.032 0.19 0.37 0.214 0.041

Stream #6 KzO 0.46 2.12 1.92 0.46 0.30 0.40 0.30 0.47

NazO 0.27 0.34 0.096 0.084 0.067 0.078 0.35 0.31

S03 0.35 1.78 1.19 0.158 0.013 0.013 0.013 <0.01

Ca <0.001 <0.001

Mn 0.003 <0.002

Fe 0.006 0.007

V <0.01 0.01

P <0.001 0.001

Ti <0.01 <0.01

KZS04 SiOz 0.14 0.02 0.019

Stream #7 Alz0 3 0.02 0.02 0.40 0.22 0.14 1.11 0.047 0.09

NazO 0.004 0.001 0.008 0.008 0.01 0.013 0.009 0.011

Cycle No. 9 10 II 12 13

Trihydrate Si02 0.17 0.27 0.41 0.32 1.13

Stream #6 K20 0.37 0.45 0.52 1.05

Na20 0.31 0.26 0.22 0.25

S03 <0.01 <0.01 <0.01 <0.01

Ca 0.003

Mn <0.002 <0.002

Fe 0.012 0.011

V <0.01 <0.01

P 0.001 0.001

Ti <0.01 <0.01

K2S0 4 SiOz <0.08 <0.02 0.43 <0.02 <0.02

4,618,480

TABLE 12-continued

Analysis of Products, Wt %

Stream #7 AI203

Na20

2.98

0.013

0.03

0.024

0.14

0.013

0.18

0.012

0.09

0.011

Stream 6 is AI(OH)).

Stream 7 is K2S04_

225 ml carbonated solution from II

By boiling with mild agitation

Solution volume maintained with H20.

Reaction in effect: 2KHC03 - K2C03 + H20 + C02

The resulting K2C03/(KHC03) solution was then

made caustic by reaction with Ca(OH)2. The conditions

and results are summarized in Table 14.

Conditions:

KHC03/(K2C03)

15 solution

Decarbonation

was maintained by adding water. After boiling for 70

10 minutes, pH of the solution rose from 8.1 to 10.6. An

82% conversion was achieved.

EXAMPLE 9

From the impurities buildup recycle tests, Example 8,

the cycles were each tested for AI, K, and S extraction.

The overall aluminum recoveries corresponded to the

single leach tests as shown in Example 8. The potassium

recovery, as K2S04, was much lower than expected, it

is believed due to the depressed solubility of K2S04 in

the presence of KOH. Recovery of K2S04 was improved

with the addition of a second water repulp stage

added in cycle 13. The sulfur extraction recovery corresponded

to the single leach tests as shown in Example 2. 20

The extraction results for the thirteen cycles are shown

in Table 13.

TABLE 14

Conditions:

K2C03/(KHC03) solution

Ca(OHh

Stoich Ca for C03 and HC03

Temperature

Time

Reactions in effect:

200 m1, decarbonated solution

13.6 g

1.10

85_90° C.

30 min, volume maintained by H20 addition

K2C03 + Ca(OHh 2KOH + CaC03

KHC03 + Ca(OHh KOH + CaC03 + H20

Results

WtlVol

Product g orml K

K2C03IKHC03 200 (75.8)

solution

Ca(OHh 13.6

Causticized solutionI 197

Residue 18.2

Conversion

Assay, gil K Distr ofKHC03

Ca HC03- C03= OH- pH % to K2C03 %

18.0 41.2 0.0 10.6 100.0

(54.1) (45.9) 0.0

100.0

0.0 10.8 21.8 78

ISoin sp gr = 1.10, % KOH = 71.9/1100 X 100 = 6.5% KOH.

TABLE 13

Cycle Extraction, %

No. Al K S

I 92.5 95.7 96.8

2 92.7 69.1 91.1

3 92.0 28.2 81.7

4 90.6 69.0 90.2

5 93.3 66.7 89.9

6 90.3 11.6 79.4

7 92.4 45.0 85.2

8 92.0 59.5 88.8

9 90.8 73.6 92.6

10 91.6 45.8 85.6

11 92.0 66.4 90.1

12 91.9 70.8 91.4

13 ~ .2±:L 96.1

Avg 91.9 64.7 89.1

EXAMPLE 10

Tests were conducted of KOH regeneration utilizing

225 ml of a process K2S04 solution which had previously

been sulfidized and carbonated to contain 75.3%

K, 0.4% S=, 8.31% S04=, 94.2% HC03-and 4.3%

C03=. The solution was boiled to convert KHC03 to

K2C03 and water by driving off C02. Solution volume

EXAMPLE 11

Initial pyrometallurgical conversion of K2S04 to

K2C03/KOH was attempted in a tube furnace. The

tube furnace consisted of a one-inch diameter quartz

50 tube surrounded by an electrically heated furnace. The

sample was placed in a silica "boat" inside the quartz

tube with a 10-20 gram mass being typical. Reaction

gases, introduced at one end, flowed over the sample

and exhausted at the other. Temperature was monitored

55 above the boat and at the tube exit.

Reagent grade K2S04 was the starting material for all

tests. Time, temperature, and reaction gas composition

were monitored on each of the six tests conducted.

Of the six tests, Tests 3, 4, and 6 showed appreciable

60 conversion to K2C03 yielding final products assaying

38.5%, 98.6%, and 78.8% K2C03, respectively. Very

little reaction took place in Test 1 and Tests 2 and 5

completely volatilized.

Tests 3 and 4 used a reaction gas consisting of 66%

65 CO, 33% N2' Test 3 lasted 15 minutes and Test 4 lasted

30 minutes. Both were run at 900· C. The reaction gas

for Test 6 was 66% H2 and 33% N2 at 850· C.

Table 15 summarizes these data.

TABLE 15

4,618,480

Tube Furnace Tests

Test Temp Time Reaction Gas, %

No. 'C. min CO H2 N2 Results

850 60 66 33 16.3% Wtloss

X-ray analysis: 10% KOH

Majority K2S04

Z 900 60 66 33 Sample completely volatilized

3 900 15 66 33 IZ.4% wt loss

Final product analysis: 51.0% K2S04

38.5% K2C03

4 900 30 66 33 Small amount of sample recovered

Final product analysis: 1.0Z% K2S04

98.6% K2C03

5 800 15 0 66 33 Sample completely volatilized

6 850 15 0 66 33 39% wtloss

Final product analysis: 10.7% K2S04

78.8% K2C03

EXAMPLE 12 the walls had 19.8% KZC03· This indicates that the

o higher temperature of the walls, 750· C., helped the

Following the tube furnace test series, a series of five 2 reaction. In actual practice, a high bed temperature

fluid-bed reactor tests were made to convert KZS04 to (760· C.-800· C.) would be necessary to give higher

KZC03 based upon DTA-TGA data and therrnody- conversions.

namic computer models. In Test 4 a 100% KZC03 bed was used because (1) it

The reactor used was a 4-inch diameter electrically should allow higher temperatures (780· C.-800· C.)

heated unit constructed of 316 stainless steel. A pre- 25 before any major fusion problems became apparent, and

heater for the reaction gases was added for Tests 3-5. (2) it would simulate more closely an actual bed. With a

Test 1 was at 675· C. with a reaction gas consisting of carbonate bed it would, hopefully, tend to agglomerate

10% Hz, 20% Nz. 65% COz. 5% HzO. The bed temper- to the carbonate provided unreacted sulfate did not

ature was increased to 760· C. with the reactor walls at build up.

850' C. A considerable amount of HzS was detected in 30 The reaction gases for Test 4 were 25% Hz, 20% Nz,

the off-gas indicating the conversion was taking place. 50% COz, and 5% HzO. The equilibrium wall tempera-

As the test continued, the bed temperature decreased ture was 860· C., bed temperature was 780· C. and

even though the wall temperature was the same. At 4 reaction gas was at 680· C. A feeder was in place to

hours, the reactor was shut down and the bed was ob- 35 slowly feed KZS04 to the bed. A target rate was 3 _

served to have fused to the walls of the reactor thereby grams/minute.

_ causing the temperature decrease. Total conversion was Test 4 was quite successful in the conversion of the

45.9%. K2S04 to K2C03 with 86.1% being converted. Prob-

Test 3 used the same reaction gas as in Test 2. The lems with the feeder allowed only 140 g of K to be

starting bed was 100% K2S04. A preheater to heat the 40 added over a 270-minute period and the·86.1% converreaction

gases was in place in an attempt to lower the sion was only on this small amount.

temperature differential between the walls and bed. The Test 5 was an attempt to repeat the Run 4 work with

bed temperature at equilibrium was 720· C. with the the feeder problems corrected. Unfortunately, a large

walls at 750· C. and the reaction gases at 507· C. The amount of unreacted sulfate, allowed to build up in the

test lasted three hours as a pressure buildup from the bed, fused as the reaction took place. Total conversion

bed caking was detected. Some fusion to the reactor 45 was 80.3% (KZS04-K2C03)'

walls was present. Total conversion for the bed material Table 16 summarizes the fluid-bed reactor tests.

was very low (1.1% KZC03) but the material fused to

TABLE 16

Fluid bed Reactor - Conversion of K2S04 to K2C03

Equilibrium Temp, 'c.

Pre- Gas Composition

Test heated Feed % Off-gas, % (Equilibrium)

No. Bed Wall Gas, 'C. H2 N2 CO2 H2O H2 N2 CO2 CO H2S

675 746 Not used 10 ZO 65 5 8.1 Z1.0 61.2 0.0 8 ppm

2 760 850 Not Used 20 20 55 5 9.6 Z4.9 53.Z 2.5 0.37%

Fused reactor

718 750 507 20 20 55 12.0 24.9 56.0 3.4 0.38%

Sample

Initial Bed

4 hr Bed

Initial Bed

15 min Bed

30 min Bed

60 min Bed

120 min Bed

180 min Bed

240 min

Final Bed

Final Wall

Initial Bed

15 min Bed

30 min Bed

60 min Bed

120 min Bed

180 min Bed

Assay, %

100 0.0

100 0.\

100 0.0

99.5 0.184

99.9 0.53

98.7 0.88

99.9 0.51

99.6 0.46

45.4

49.0 41.5

100 0

0.01

99.5 0.01

99.9 0.01

98.9 0.35

98.5 1.1

Conversion

S04 C03

45.9

4,618,480

TABLE 16-continued

Fluid bed Reactor - Conversion of K2S04 to K2C03

Test

No.

Equilibrium Temp. DC.

Preheated

Feed %

Bed Wall Gas."c. H2 N2 C02 H20

Gas Composition

Off-gas. % (Equilibrium)

Sample

Assay. %

Conversion

25 20 50 5 12.6 23.7 44.6 12.8 320 ppm

Run 4 had a continuous feed

of K2S04 to a K2C03 bed

19.8 (19.8)1

4 780 860

740 830

680

680 25 20 50 6.8 24.5 47.7 22.2 1140 ppm

180 min

Final Wall

Initial Bed

30 min Bed

60 min Bed

120 min Bed

150 min Bed

180 min Bed

210 min Bed

270 min Bed

Initial Bed

15 min Bed

60 min Bed

90 min Bed

120 min Bed

150 min Bed

180 min Bed

210 min Bed

240 min Bed

260 min

Final Bed

(80.2)1

1.46

3.34

5.20

5.28

13.3

3.82

6.81

2.30

1.66

3.51

14.8

20.3

20.5

20.3

15.9

8.01

4.59

4.74

100

98.5

91.0

91.7

86.7

96.1

90.3

92.8

95.0

91.2

81.1

77.6

75.3

76.6

81.8

86.8

91.0

91.8

23.1

27.4

86.1

oooo

14.8

61.9

80.3

IEstimate based on the K2C03 assay.

% Conversion

K2S04KHC02

Amount Assay. % or gil

Sample g or ml K Ca S04

Feed liquor 1200 (64.6) (sal'd) (79.4)

Filtrate 994 62.3 0.87 13.8

65 Precipitate! 119.4 6.42 26.4 56.7

121.1% cake moisture· filter rate 52 gallonslhr ft2.

2Based on sulfate assays.

30 KCOOH, with the subsequent conversion of the

KCOOH product to potassium carbonate, K2C03.

The conversion of K2S04 to KCOOH was carried

out in an autoclave. The feed liquor comprised a 1200

cc solution comprising 120 gramslliter K2S04. The

35 solution was contacted' with 72 grams of Ca(OH)z, approximately

1.5 times the stoichiometric amount. A

large amount of CO excess was added at a flow rate

equalling 4 liters per minute. The natural pressure of the

autoclave was approximately 220 psi, with CO added to

40 increase the pressure to approximately 500 psi. The

temperature of the test was conducted at 220· C. The

test ran for 15 minutes at the specified temperature and

pressure. An 88.2 percent conversion of K2S04 to

KCOOH occurred. Table 18 summarizes the results of

45 the test.

In the second part of the testing, for the conversion of

KCOOH to K2C03, the reaction was carried out in a

furnace with atmospheric oxygen as the only oxidizing

agent. Potassium formate crystals from the initial testing

50 were produced by evaporation of the liquor. These

formate crystals were then tested for the evaluation of

the conversion of potassium formate to potassium carbonate.

At 200· C., the crystals melted, however, no

reaction was noted over a 2-hour period (the melting

55 point of KCOOH is approximately 170· C.). The temperature

was then increased to 400· C. Crystals were

produced at this temperature. These crystals were assayed

at 77 percent K2C03.

TABLE 18

99.54

Conversion

K2C032KOH

TABLE 17

Conversion of Potassium Carbonate

to Potassium Hydroxide

Test Description and Results

EXAMPLE 14

A test was performed to evaluate the conversion of

potassium sulfate, K2S04, to potassium formate,

Filtrate! 848 34.9 8.40 88.1 (2.10)3

Precipi- 100.3 0.0 1.622 0.366 38.5

tate (96.0) balance. %

IThe amount of filtrate contained in the wash was calculated from the wash liquor 60

analysis and this value added to the filtrate volume.

2A major portion of the precipitate was found by analysis to be He03-.

3Calculated value.

4Based on K2CO) added and K2CO) in the residue.

Amount Analysis. % or gil

Sample g or ml OH C03 K Ca

Reaction: K2C03 + Ca(OHh - 2KOH + CaC03 (insoluble)

Conditions: K2C03. g 138.2

Ca(OHh. g 74.7 (I X stoichiometric)

H20. ml 1000

Temp 30 min at ambient

60 min at 97" C. (boiling)

A 30-minute sample was taken and the pH was 11.0. This was

too low to have a significant concentration of OH- present. The

solution was taken to boiling for 60 minutes with a much higher

pH noted (13 +). The slurry was filtered and the cake washed

with water.

EXAMPLE 13

A test was undertaken to verify the following reaction

for the conversion of K2C03 to KOH:

This test was initially run at ambient temperature with

no detectable reaction after 30 minutes. It was then

heated to boiling (100· C.) for 60 minutes with 99.5%

conversion. Table 17 describes the conditions and resuIts

of this rest.

--------...T.A;B.L.E;1=9 --------- 25

Assay, % or gil

5. A process according to claim 1 further comprising

calcining said AI(OH)3 crystals to produce alumina.

6. A process according to claim 1 further comprising

crystallizing K2S04 from at least a portion of said secondary

leach to form a spent liquor.

7. A process according to claim 1 further comprising:

(e) regenerating KOH for recycle to step (a) from at

least a portion of said secondary leachate of step

(d).

8. A process according to claim 7 wherein step (e)

comprises contacting a portion of said secondary leachate

of step (d) with lime and hydrogen sulfide to form

calcium sulfate and a potassium sulfide- and potassium

hydrogen sulfide-containing liquor; carbonating said

liquor to form potassium carbonate; and causticizing

said potassium carbonates to produce potassium hydroxide.

9. A process for recovery of alumina and potassium

sulfate from alunite ore containing AI, K and S values,

comprising:

(a) contacting said ore at a temperature above 600 C.

with potassium hydroxide saturated with powsium

sulfate to which no sodium has been added to

form a potassium sulfate-saturated primary potassium

hydroxide leach liquor containing said Al

values and a primary leach residue containing said

K and S values;

(b) separating said primary leach liquor from said

primary leach residue;

step (b) to recover Al values therefrom;

(d) calcining said AI(OHh crystals to produce alumina;

(e) aqueous leaching said primary leach residue of

step (b) to form a secondary leachate containing

said K and S values; and

(f) crystallizing K2S04 from at least a portion ofsaid

secondary leachate to form a spent liquor.

10. A process according to claim 9 further comprising

desilicating the liquor of step (b) prior to step (c).

11. A process according to claim 10 further comprising

regenerating KOH for recycle to step (a) from at

least a portion of said secondary leachate of step (e).

12. A process according to claim 11 wherein said

regenerating comprises contacting a portion of said

secondary leachate of step (e) with lime and hydrogen

sulfide to form calcium sulfate and a potassium sulfideand

potassium hydrogen sulfide-containing liquor; carbonating

said liquor to form potassium carbonate; and

causticizing said potassium carbonates to produce potassium

hydroxide.

13. A process according to claim 9 further comprising

recycling at least a portion of the spent liquor of step (f)

to step (e); and controlling the level of KOH in said

second leach liquor of step (f) by recycling at least a

portion of the spent liquor to step (a).

14. A process according to claim 13 further comprising

controlling the build-up ofimpurities during step (d)

by removing said impurities from at least a portion of

said desilicated leach liquor.

15. A method of producing Ah03 and K2S04 from

alunite ore using Cao as the only make-up reagent comprising:

(a) contacting said ore at a temperature above 600 C.

with a K2S04-saturated KOH to which no Na has

been added to form a K2S04-saturated first KOH

leach liquor containing said AI values and a first

leach residue containing said K and S values;

4,618,480

% Conversion

KZS04 -> KHCOz:

97.4%

KHCOz -> KzC03:

22.1%

KZS04 -> KOH: 21.5%

28.7

2.70 1.10 25.7

57.6 2.0

94.7

(K)

0.138

Balance, %

Filtrate

Precip.

Sample

Although the foregoing invention has been described

in some detail by way of illustration and example for

purposes of clarity of understanding, it will be obvious 35

that certain changes and modifications may be practiced

within the scope of the invention, as limited only by the

scope of the appended claims.

What is claimed is: 40

1. A process for recovering aluminum and potassium

values from alunite ore comprising:

(a) contacting said ore at a temperature above 600 C.

with potassium hydroxide saturated with potassium

sulfate to which no Na has been added to 45

form a potassium sulfate-saturated primary potassium

hydroxide leach liquor containing said Al

values and a primary leach residue containing said

K values;

(b) separating said primary leach liquor from said 50

primary leach residue;

step (b) to recover the Al values therefrom; and

(d) aqueous leaching said primary leach residue of

step (b) to form a secondary leachate containing 55

said K values.

2. A process according to claim 1 further comprising

desilicating the liquor of step (b) prior to precipitating

said AI(OH)3.

3. A process according to claim 2 wherein said desili- 60

cation is by contacting said primary leach liquor with

CaO at a temperature of from about 1800 to about 30

2000 C. and for a retention time of from about 5 to about

minutes.

4. A process according to claim 3 wherein said CaO 65

is in an aqueous solution as Ca(OHh in an amount of

from about 12 to about 20 grams per liter of said first

leachate.

EXAMPLE 15

Testing was done to evaluate a direct, one-step hydrometallurgical

conversion of K2S04 to KOH. The

first reaction (involved contacting potassium sulfate 5

with hydrated lime under an atmosphere of carbon

monoxide to produce potassium formate and gypsum.

This reaction took place at 500 psi and 2250 C. The

system was allowed to react for 30 minutes with a CO

sparge of 4 liters per minute. The next reaction involved 10

oxidizing the potassium formate produced in the first

reaction with oxygen at 500 psi and 225 0 C. to yield

potassium carbonate. The formed potassium carbonate

immediately reacted with excess hydrated lime to produce

potassium hydroxide. 15

The potassium sulfate and calcium hydroxide were

added as a 25 percent solids, by weight, slurry to a

pressure autoclave. The autoclave was heated to 2250

C., then a·CO atmosphere was sparged through the 20

system at 500 psi for 30 minutes. Oxygen was then

sparged through the unit for 30 minutes under the same

conditions. The results are shown in Table 19.

4,618,480

(ii) treating the sulfide products of step (i) with water

and carbon dioxide to form potassium carbonate,

potassium bicarbonate and hydrogen sulfide;

(iii) heating the potassium bicarbonate of step (ii) in

the presence ofwater to form potassium carbonates

and carbon dioxide;

(iv) treating said potassium carbonates with lime to

generate potassium hydroxide, carbon dioxide and

calcium carbonate, and separating said potassium

hydroxide.

20. The process ofclaim 19 in which hydrogen sulfide

from step (ii) is recycled to step (i).

21. The process of claim 19 in which carbon dioxide

from step (iii) is recycled to step (ii).

22. The process of claim 19 in which carbon dioxide

from step (iv) is recycled to step (ii).

23. The process of claim 19 in which calcium carbonate

from step (iv) is decomposed to lime and carbon

dioxide.

24. The process of claim 19, in which formed lime is

recycled to step (i).

25. The process of claim 23 in which formed carbon

dioxide is recycled to step (i).

26. The improvement according to claim 16 in which

said potassium hydroxide is generated for recycle to

step (a), the improvement further comprising:

(i) treating a portion of said potassium sulfate solution

of step (c) with lime and carbon monoxide under

pressure to form potassium formate in solution and

a calcium sulfate precipitate;

(ii) performing a liquid/solid separation of the products

of step (i);

(iii) crystallizing potassium formate from the liquid of

step (ii) and separating the crystals from the mother

liquor;

(iv) oxidizing the crystals of step (iii) to potassium

carbonate;

(v) treating potassium carbonate of step (iv) with lime

to generate potassium hydroxide, carbon dioxide

and calcium carbonate, and separating said potassium

hydroxide.

27. The improvement according to claim 16 in which

said potassium hydroxide is generated for recycle to

step (a), the improvement further comprising treating a

portion of said potassium sulfate of step (c) with lime

and carbon monoxide at elevated temperatures and

pressures sufficient to form potassium formate and calcium

sulfate; adding an oxidizing agent to convert at

least a portion of the potassium formate to the carbonate,

and form carbon dioxide and water; reacting the

formed carbonate to react with lime to produce potassium

hydroxide and calcium carbonate, and separating

said hydroxide.

28. The process of claim 27, in which said added

oxidizing agent is oxygen.

29. The process of claim 27, in which lime is added

for treatment of said potassium sulfate solution in an

amount in excess of that needed to react with all the

potassium sulfate present, and no lime is added in con-

60 junction with said oxidizing agent.

30. The improvement according to claim 16 in which

said potassium hydroxide is generated for recycle to

step (a), the improvement further comprising contacting

a portion of said potassium sulfate solution of step

for a time sufficient to produce potassium carbonate and

hydrogen sulfide and contacting said potassium carbonate

with an aqueous Ca(OHh solution to produce a

(b) separating said first leach liquor from said first

leach residue;

separating the precipitated silicates from said liquor;

(d) precipitating Al(OH)J crystals from the desilicated

liquor of step (c) to recover the Al values

therefrom;

(e) calcining said Al(OH)J crystals to produce alumina;

(f) leaching of said first leach residue of step (b) with

an aqueous solution to form a second leach liquor

containing said K and S values;

(g) crystallizing K2S04 from at least a portion of said

second leach liquor to form a spent liquid; 15

(h) regenerating KOH for recycle to step (a) by contacting

a portion of said secondary leachate of step

(f) with lime and hydrogen sulfide to form calcium

sulfate and a potassium sulfide- and potassium hydrogen

sulfide-containing liquor; carbonating said 20

liquor to form potassium carbonate; and causticizing

said potassium carbonates to produce potassium

hydroxide;

(i) recycling at least a portion of the spent liquor of 25

step (g) to step (f);

U) controlling the level ofKOH in said second leach

liquor of step (g) by recycling to step (a) at least

one of the streams selected from the group consisting

of (1) at least a portion of the spent liquor of 30

step (g); and (2) at least a portion of said second

leach liquor of step (f);

(k) controlling the build-up of impurities in the crystallization

of step (d) by removing said impurities

from at least a portion of said desilicated liquor; 35

(1) heating said CaC03 of step (h) to form CaO and

C02; and .

(m) recycling said CaO to step (c) and step (h).

16. In a process for recovering aluminum values and

potassium sulfate from alunite ore containing aluminum, 40

sulfur and potassium values wherein the ore is treated

with a caustic leach to solubilize aluminum values into

the leachate, the improvement comprising:

(a) leaching with potassium sulfate-saturated potassium

hy(lroxide to which no sodium has been 45

added to form an aluminum-containing primary

leachate and a primary residue containing said

potassium and sulfur values;

(b) separating said leachate from said residue;

said residue to form a potassium sulfate solution

and recovering potassium sulfate therefrom; and

(d) recovering aluminum from said leachate.

17. The improvement according to claim 16 in which

the ore further contains silicon values, the process fur- 55

ther comprising desilicating said leachate.

18. The improvement according to claim 17 in which

said desilicating comprises contacting said leachate

with lime; recovering aluminum from said solution. as

aluminum hydroxide crystals; and calcining said crystals

to produce alumina.

19. The improvement according to claim 16 in which

potassium hydroxide is generated for recycle to step (a),

the improvement further comprising:

(i) treating a portion of said potassium sulfate solution 65

of step (c) with lime and hydrogen sulfide to form

calcium sulfate, potass:um sulfide, potassium hydrogen

sulfide and water;

* * * * *

portion of said potassium sulfate solution of step (c)

with barium oxide to form potassium hydroxide and

barium sulfate; and separating said potassium hydroxide.

36. The process of claim 35 in which barium oxide is

regenerated from barium sulfate and recycled to form

additional potassium hydroxide.

37. The process of claim 36 in which said regenera-

10 tion comprises the steps of reducing said barium sulfate

with coal to barium sulfide; reacting said barium sulfide

with carbon dioxide to form barium carbonate and hydrogen

sulfide; and reducing said barium carbonate to

barium oxide.

4,618,480

KOH solution and a CaC03 precipitate and separating

the KOH solution for recycle to step (a).

31. The process according to claim 30, in which said

contacting occurs at temperatures of from about 600· C.

to about 1000· C. and in the presence of H20.

32. The process of claim 30 in which the reducing

agent comprises coal.

33. The process of claim 30 in which the reducing

agent comprises a mixture of hydrogen gas and carbon

dioxide.

34. The process of claim 30 in which the reducing

agent comprises carbon monoxide.

35. The improvement according to claim 16 in which

said potassium hydroxide is generated for recycle to

step (a), the improvement further comprising treating a 15