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]
m~
[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
19
l' -I
I
I
I
I
I
I
LIM E 8
REGENERATION
u.s. Patent Oct 21, 1986 Sheet 2 of4 4,618,480
,...., a
N
: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)
3
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
2
~ 0
N
~ K2SCS.L)
"(
f)
N -I
~
~
5
u.s. Patent Oct 21, 1986 Sheet 4 of4
FIGURE 4
4,618,480
ALUMINUM EXTRACTION vs TIME
zo;:::
~
X""""
IJ.I
90
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
1
TECHNICAL FIELD
2
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
3
SUMMARY OF THE INVENTION
4
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
6
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)
5
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:
8
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
7
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
9
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:
10
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
12
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)
35
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
11
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:
14
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.
13
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
15
TABLE I-continued
Element
Ga
Total organic carbon (TOC)
%
0.002
1.81
4,618,480
16
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
DE
F Raw ore
GH
I Raw ore
J
K
L Raw ore
MN
(pH 10.0)
(pH 10.0)
10% 1.4
18% 2.8
10% 1.4
18% 2.8
90
70
90
90
130
150
5
10
20
60
120
60
5
10
40
5
15
40
2
5
10
2
5
10
11.4
0.0
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
2
4
6
7
9
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
100
100
100
60
60
60
60
80
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
2
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
0.031
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
17
4,618,480
18
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.
19
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
5
A/K Final Filtrate
Target Ratio
0.7
20
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
om
0.005
Ore
Major2
10
I
0.1
om
4.0
0.3
0.002
0.01
0.001
0.1
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%.
30
25
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
21
4,618,480
22
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
23
4,618,480
24
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.
25
TABLE 15
4,618,480
26
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
oo
45.9
27
4,618,480
28
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
oo
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
30
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;
(c) precipitating Al(OHh crystals from the liquor of
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;
30
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
(K)
0.138
80
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;
(c) precipitating AI(OH)3 crystals from the liquor of
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.
29
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.
5
10
4,618,480
32
(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
(c) with a reducing agent at elevated temperature and
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
31
(b) separating said first leach liquor from said first
leach residue;
(c) desilicating said separated first leach liquor and
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;
(c) solubilizing said potassium and sulfur values in 50
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;
* * * * *
5
34
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
33
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
20
25
30
35
45
50
55
60
65