13 Claims, 1 Drawing Figure
A sequential flotation process for the recovery of highgrade
concentrates of copper, lead and cobalt-nickel
from sulfide ores is provided. A primary grind ore pulp
is conditioned with S02 as H2S03 under intense aeration,
and the conditioned pulp subjected to sequential
flotation, with regrinding and conditioning of a copper
rougher concentrate obtained in the first flotation step
for copper.
3,309,029 3/1967 Frame 209/166
4,283,017 8/1981 Coale et aJ. 209/167
4,351,668 9/1982 Stephenson et aJ. 75/2
4,387,034 6/1983 Unger et aJ. 209/166
FOREIGN PATENT DOCUMENTS
54-26482 9/1979 Japan 209/167
711887 3/1971 South Africa.
Primary Examiner-Bernard Nozick
Attorney, Agent, or Firm-Holman & Stern
United States Patent [19]
Shaw et al.
[54] SEQUENTIAL FLOTATION OF SULFIDE
ORES
[75] Inventors: Douglas R. Shaw; John F. Spisak,
both of Arvada; Jerome P. Downey,
Parker; Gary E. Butts, Arvada, all of
Colo.
[73] Assignee: Anschutz Mining Corporation,
Denver, Colo.
[21] Appl. No.: 466,837
[22] Filed: Feb. 16, 1983
[51] Int. CI.3 B03B 1/00
[52] U.S. CI. 209/9; 209/167;
252/61; 75/2
[58] Field of Search 209/166, 167, 4, 9;
241/11; 252/61; 75/2
[56] References Cited
U.S. PATENT DOCUMENTS
2,399,845 5/1946 Allen et aJ. 209/167
ORE
[57]
[II] Patent Number:
[45] Date of Patent:
ABSTRACT
4,460,459
Jul. 17, 1984
Pb
CONCENTRATE
u.s. Patent JuI. 17, 1984 4,460,459
ORE
• FINAL
TAILINGS
Co-Ni1st
CLEANER
Co-Ni3"d
CLEANER
Co-Ni 2"d
CLEANER
Co-Ni
ROUGHER
Co-Ni
CONCENTRATE
Pb2nd
CLEANER
Pb 1st
CLEANER
Cu
CONCENTRATE
Pb
CONCENTRATE
4,460,459
1
SEQUENTIAL FLOTATION OF SULFIDE ORES
BACKGROUND OF THE INVENTION
Sulfide ores of the type common to the lead belt areas 5
of southeastern Missouri typically have a valuable mineral
content of copper, lead and cobalt-nickel. Characteristically,
much of the cobalt-nickel content is lost in
the conventional treatment of these ores for recovery of
the copper and lead content, and cobalt-nickel is mainly 10
recovered as a low-yield by-product.
The sequential flotation method of the invention applied
to such ores permits the recovery of high-yield
concentrates of copper, lead and cobalt-nickel. While
various selective flotation methods have been applied to 15
complex ores containing copper, lead and zinc mineral
suites, with successful recovery·of zinc, these ores' are
mineralogically very distinct from the ore starting material
of the present invention, and the prior art has not
succeeded in the practical- application of sequential 20
flotation to the subject sulfide ores.
SUMMARY OF THE INVENTION
The invention provides a sequential flotation process
for the primary recovery ofhigh-grade concentrates of 25
copper, lead and cobalt-nickel from sulfide ,ores of the
type common to the Missouri .lead belt area of North
America. Concentrates of copper,' lead' and· combined
cobalt and nickel are ,separately recovered,in that order
by the chemical control. and manipulation of the flota- 30
tion rates of the copper,lead, cobalt-nickel and iron
sulfide minerals present' in the ore in a conventional
sequential flotation system'comprising,a main flotation
circuit for each of the product concentrates. Broadl.y,
according to the process, ground oxe pulp is condi- 35
tioned with sulfur dioxide and intensely aerated prior to
copper flotation; the copper rougher concentrate fl'0tn
the copper flotation circuit is relatively fin,ely reground
and conditioned with sulfur dioxide prior to cleaning.
Preferably, the main copper circuit tailings are fouted40
to the lead and cObalt-nickel· flotation circuits in an
open-circuit manner.
BRIEF DESCRIPTION OF THE DRAWING
The sole FIGURE is a flowsheet of a continuous 45
sequential flotation process according to the inveri.tion.
DETAILED DESCRIPTION OF THE
INVENTION
The process of the invention is specifically directed 50
.to the recovery of separate concentrates of copper, lead
and cobalt-nickel from siegenite-bearing oreS of the
type common to deposits broadly classified as Mississippi
Valley-type deposits. The ores are characterized
by sulfide mineral suites typically occurring as siegenite 55
or linnaeite (cobalt-nkkel) with chalcopyrite (Cu), galena
(Pb), and usually marcasite (Fe), in a carbonate
matrix such as dolomite or calcite, and are exemplified
by the siegenite-bearing ores of southeastern Missouri
and the viburnam trend ore bodies of the new lead belt. 60
The are starting material of the present process is
ground to sufficiently liberate sulfide minerals for subsequent
flotation. In general, a primary grind fineness
(ball mill) of from about 65% to about 75% passing 200
mesh (Tyler) is suitable; however, the ease of sulfide 65
liberation with relatively coarse grinding may permit
the use of a primary grind product of 60% or less passing
200 mesh, depending on the ore characteristics. The
2
flotation characteristics of the primary grind product
are also dependent upon the grinding medium employed,
and the fineness of the grind is accordingly
adjusted .to autogenous, semi-autogenous, pebble or
other milling procedures, as necessary.
After grinding, the primary grind pulp is conditioned
to depress lead, iron and cobalt-nickel sulfides by addition
of sulfur dioxide, preferably in the form of sulfurous
acid, and aerated to enhance the promotion and
flotation rate of copper. Preferably, S02 is added in an
amount of from about 1 to about 5 Ibs S02 per ton of
pulp; the amount will vary, however, depending on the
flotation conditions and characteristics of the flotation
pulp. If natural air is employed, aeration at a rate of
about 3 to 5 cu ft/min per cubic foot of pulp generally
will satisfactorily promote copper. Generally, the pulp
is aera~ed substantially concurrently with S02 addition,
although the sequence ofS02 addition and aeration may
be varied within broad limits with satisfactory results,
depending on actual conditions.
The conditioned pulp is then routed to a flotation
systeniof the type schematid111y illustrated in the sole
Figure, comprising three' main flotation circuits for
tecovery of copper, lead' and cobalt~nickel, respectively.
(Generally', the recovery of iron present in the
subject ore bodies is not economically feasible.) Each of
the circuits includes successive concentration and separation
stages' comprising a roughing stage wherein a
rougher concentrate is recovered, and a plurality of
cleaning .stages, wherein the rougher qoncentrate is
up-graded. Tailing products from each of the circuits
are routed to the' next circuit for addjtional mineral
recovery. .
Flotation of coppeds effected in the copper flotation
circuit at a slightly acidic pulp pH of about 6.5 to 6.8,
the pH being governed by the quantity of sulfur dioxide
(S02) used during conditioning and aeration. A collector
selective for copper in an acidic medium is employed,
such as ethyl isopropyl thionocarbamate. The
pulp is frothed for a period of time which maximizes
copper recovery with minimal II}isplacement oHead or
cobalt-nickel; typically, froth times of two to four minutes
are adequate. The copper rougher concentrate is
then collected, and the copper rougher tailing product
is routed to the lead flotation circuit.
The copper rougher concentrate is finely reground
prior to cleaning to further liberate cobalt-nickel minerals
present and improve their rejection (see Table 1).
While regrinding does not generally affect lead recovery,
the rougher concentrate should not be reground so
finely that the flotation properties of copper are adversely
affected. In general, regrinding power requirements
of 10 kwhr/ton to about 50 kwhr/ton, preferably
from about20 to 30 kwhr/ton are suitable. The regound
concentrate is then conditioned with S02, again advantageously
as sulfurous acid, to depress liberated cobaltnickel
sulfides, usually fit amounts of from about 0.05
lbs. to about 1.5 lbs. S02 per ton of reground pulp. The
reground concentrate is then cleaned in a conventional
way, for example, by addition of collector S02 and
sodium dichromate. Preferably, the first copper cleaner
tailings are combined with the copper rougher tailing
product and routed to the lead flotation circuit, rather
than recycling the cleaner tailings to the copper
rougher as is customary, as this promotes better lead
and cobalt-nickel recovery. The copper cleaner product
is cleaned one or more times, as desired, and a highEXAMPLES
Tables 2-4 summarize data on reagent suites and
operational conditions for three pilot plant runs accord35
ing to this invention.
4,460,459
Copper Concentrate
Cu Regrind, Assay, % Distribution, %
kwhr/ton Cu Ph Co Ph Co
IA com""ral;" Ial wilhoul a copper circuil relrind wu nol conducted on Ihis 15
_pie.
3
purity copper concentrate, typically containing in excess
of 85% of original copper values, is recovered.
TABLE 1
)
0 28 3.4 0.S7 7.S 10.0
Sample 2
30 31 6.S 0.18 11.7 2.1
)
0 26 4.1 O.SS 9.4 12.9
Sample 3 14 3
30
1 4.3 0.34 8.8 7.4
29 4.S O.IS 7.8 2.9
)
81 2S S.O O.IS 18.S S.9
Sample S
13 32 2.2 0.31 8.6 3.4
4
purity cobalt-nickel concentrate containing up to about
92% of the values originally present.
Numerous variations within the scope of the invention
will be apparent. Sulfur dioxide, a strong reducing
5 agent, is a key reagent, providing selectivity control
throughout the system. In the highly reduced environment
provided by S02, intense aeration depresses lead
and any iron sulfides present by selective surface oxidation,
and also promotes copper and enhances its flota-
10 tion rate. Various copper collectors in addition to the
ethyl isopropyl thionocarbamate mentioned are useful,
with the caveat that they retain selectivity in the acid
environment present; copper collectors such as xanthates
and dithiophosphates, for example, may promote
considerable lead flotation with the copper. Generally,
known collectors, frothers and other reagents are contemplated
for use in the lead, copper and cobalt-nickel
Lead and cobalt-nickel are recovered as concentrates flotation circuits. Froth times in all circuits are varied as
from the respective flotation circuits in conventional necessary to maximize recoveries. The use of lime to
fashion. In an exemplary embodiment, lead is recovered 20 adjust the pH in the cobalt-nickel flotation circuit is not
by flotation after adjustment of the pH of the pulp to recommended, as this tends to increase viscosity and
about 8.5 to 9 and after depression of the cobalt-nickel interfere with flotation.
sulfides present by addition of sodium cyanide in an The concentration conditions of the flotation circuits
amount of from about 0.25 to 0.375 lb/ton, followed by may be adjusted to the prevailing circumstances within
collector addition and frothing for about 3 to 5 minutes. 25 broad limits. Generally, at least three cleaning stages
(While greater amounts of cyanide tend to improve ' are employed in each circuit, typically in a conventional
cobalt-nickel rejection in the lead circuit, they also tend countercurrent flow pattern. Tailings are cycled as
to severely depress cobalt-nickel and interfere with necessary to optimize recovery of a particular mineral.
subsequent flotation.) Similarly, cobalt-nickel is recov- Additional adaptations within the scope of the invenerable
by flotation after addition of copper sulfate, 30 tion will be apparent to those skilled in the art.
which activates cobalt-nickel and complexes with excess
cyanide present. After a cobalt-nickel rougher
froth time of about 8 minutes or more to maximize
cobalt-nickel recovery, the cobalt-nickel rougher concentrate
is recovered and cleaned to provide a high-
ExampleI, (Table 2),Cycle test CT-3, Sample 2
TABLE 2
Cycle Test CT-3 Test Conditions
Pilot Plant Sample 2
State
Reagents Added, PoundslTon __T.:.;i:;::m",eL...'M=in::,:u:;;tes::::-__ Pulp
AX-3433 MIBC4 Grind Cond Froth pH
10 6.5
om 1 I.S 6.5
1 I.S 6.5
0.10 20
0.005 I 4 6.5
0.05 1 3 6.5
0.04 I 2 6.5
1.0 0.30 10 9.0
0.02 0.015 0.01 I
Cu regrind Rougher Cleaners
5" X 7" pebble mill 1000 g 0,1 250 g 0-1
72
Primary grind
Aeralion
Cu rougher (I)
(2)
Cu regrind
Cu 1st cleaner
Cu 2nd cleaner
Cu 3rd cleaner
Ph condilioning
Pb rougher
Stage
Equipment
Speed (rpm)
% solids
I.S
0.7S
0.016
0.10
0.20 0.008
0.10 0.008
0.10
0.10
Primary grind
5" X 12" batch mill
52
65
0.20 20
Pb 1st cleaner
Pb 2nd cleaner
Pb 3rd cleaner
Pb 4th cleaner
Co, Ni conditioning
Co, Ni rougher (I)
(2)
Co, Ni 1st cleaner
Co, Ni 2nd cleaner
Co, Ni 3rd cleaner
Stage
Ca(OHh
0.10
0.05
0.05
0.05
Roughers
NaCN
0.05
0.Q25
0.Q25
0.Q25
Reagents Added, PoundslTon Time, Minutes Pulp
Na2Si03 AP-2422 AX.3433 CUS04 MIBC4 Grind Cond Froth pH
0.05 0.01 I 2 9.5
0.025 I 2
0.025 1 1
0.Q25 I I
0.6 5 8.2
0.05 I 4
0.05 0.2 2 4 8.0
0.01 I 4 7.7
0.01 I 3 7.9
0.01 I 2 7.9
Co, Ni lsI cleaner Remaining cleaners
5
4,460,459
6
TABLE 2·continued
Cycle Test CT·3 Test Conditions
Pilot Plant Sample 2
Equipment 1000 g 0·1 500g 0·1 250 g 0·1
IEthyl isopropyl thionocarbamate
2Ammonium diisopropyl dithiophosphate
3Sodium isopropyl xanthate
'Methyl isobutyl carbinol
Example II (Table 3) Cycle Test CT·4, Sample 3 10
TABLE 3
Cycle Test CT-4 Test Conditions
Pilot Plant Sample 3
Reagents Added, PoundslTon
Stage
__T.=..t:::·m:::e:z.'"'M:::in:::u:::t::::es=--__ Pilip
AX·3433 MIBc4 Grind Cond Froth pH
Primary grind
Aeration
Cu rougher (I)
(2)
Cu regrind
Cu 1st cleaner (I)
(2)
Cu 2nd cleaner
Cu jrd cleaner
Pb conditioning
Pb rougher .
Stage
Equipment
Speed (rpm)
% solids
1.0
0.70
0.024
0.008
0.10
0.10 0.008
0.008
0.10
0.06
Primary grind
5" X 12" batch mill
52
65
0.2
0.1
0.05
0.04
0.8 0.3
Cu regrind
5" X 7" pebble mill
72
50
0.016
0,02 0.015
Roughers
.1000g 0·1
26
12
Cleaners
250 g 0·1
10
I
10
I
6:5
22
6.7
2 6.3
23
2.
8.5
Ca(OHh
0.05
0.Q2
0.01
0.01
Other cleaners
250 g 0·1
1100
Time, Minutes Pulp
Grind Cond Froth pH
I 2 9.5
1 2
I I
I I 9.5
5
1 4 8.0
2 4
1 4 8.0
1 3
1 2
0.6
0.2
CUS04 MIBc4
0.05
0.05
0.01
0.01
om
Co, Ni 1st cleaner
500g 0·1
1300
0.05 0.01
0.025
0.025
0.025
Reagents Added, PoundslTon
0.05
0.025
0.025
0.025
NaCN
Pb 1st cleaner
Pb 2nd cleaner
Pb 3rd cleaner
Pb 4th cleaner
Co, Ni conditioning
Co, Ni rougher (I)
(2)
Co, Ni 1st cleaner
Co, Ni 2nd cleaner
Co, Ni 3ed cleaner
Stage Roughers
Equipment 1000 g 0·1
Speed 1600
IEthyl isopropylthionocarbamate
2Ammonium diisopropyl dithiophosphate
3Sodium isopropyl xanthate
'Methyl isobutyl carbinol
spotassium amyl xanthate
Example III (Table 4) Cycle Test CT·5, Sample 5
TABLE 4
Cycle Test CT·5 Test Conditions
C Pilot Plant Sample 5
Reagents Added, PoundslTon Time, Minutes Pulp
Stage S02 M.1661 1 Na2Cr207 Ca(OHh NaCN AP·2422 AXc3~33 MIBc4 Grind Cond Froth pH
Primary grind 1.0 0.2 26
Aeration 0.80 10 6.
Cu rougher (1) 0.024 0.01 1 2
(2) 0.008 I 2
Cu regrind 0.1 0.1 17
Cu 1st cleaner (I) 0.06 0.016 om 1 2 6.
(2) 0.008 1 3
Cu 2nd cleaner 0.12 0.05 I 3.5 6.
eu 3rd cleaner 0.06 0.04 1 2.5 6.
Pb conditioning 0.5 0.3 10 8.
Pb rougher 0.02 0.015 0.01 I 8.
Stage Primary grind Regrind Rougher Cleaners
Equipment 5" X 12" batch mill 5" X 7" pebble mill 1000 g 0·1 250g 0·1
Speed (rpm) 52 72 1800 1200
% solids 65
Reagents Added, PoundslTon Time, Minutes Pulp
7
4,460,459
8
TABLE 4-continued
Cycle Test CT-5 Test Conditions
Pilot Plant Sample 5
Pb Ist cleaner
Pb 2nd cleaner
Pb 3rd cleaner
Pb 4th cleaner
Co, Ni conditioning
Co. Ni rougher (I)
(2)
Co. Ni 1st cleaner
Co, Ni 2nd cleaner
Co, Ni 3rd cleaner
Stage
Equipment
Speed (rpm)
Ca(OHh
0.10
0.05
0.05
0.05
Rougher
lOOOgD-1
1800
NaCN
0.05
0.025
0.025
0.025
0.05
0.025
0.025
0.D25
0.01
0.5
0.05
0.05 0.2
0.01
0.01
0.01
Co, Ni 1st cleaner
500 g D-I
1500
MIBC4 Grind Cond Froth pH
I 2 9.5
I 2
I I
I I 9.5
5 8.5
I 4 8.5
2 4
I 4 8.0
I 3
I 2
Remaining cleaners
250 g D-I
1200
1Ethyl isopropyl thionocarbamate
2Ammonium diisopropyl dithiophosphate
3Sodium isopropyl xanthate
-4Methyl isobutyl carbinol
sPolUSium amyl xanthate
Example IV-Table 5 summarizes the results obtained
from cycle testing according to Examples I, II and III.
As much as 91% of the copper, 85% of the lead and
92% of the cobalt and nickel values were recovered in 25
their respective concentrates. Cycle tests were not con·
ducted on Samples 1 and 4. A primary grind of 60 to
70% passing 200 mesh was employed. Thickening and
filtration rates of the products were judged adequate to
good.
effecting flotation of the copper and separating a
copper rougher concentrate from a copper rougher
tailing product;
regrinding the copper rougher concentrate to liberate
lead and cobalt-nickel minerals and conditioning
the reground concentrate with S02;
cleaning the reground conditioned rougher concentrate
and separating a first copper cleaner concentrate
from a first copper cleaner tailing product;
TABLE 5
Weight Assays, % Distribution, %
Product % Cu Pb Co Ni Cu Pb Co Ni
Sample No.2
Cu conc 2.51 28.6 4.68 0.19 0.27 89.0 11.6 3.3 3.0
Pb conc 1.01 0.84 79.2 0.14 0.18 1.0 78.9 1.0 0.8
Co-Ni COnc 3.24 1.16 1.05 3.80 5.85 4.7 3.4 86.1 82.5
Head (calc) 0.81 1.01 0.143 0.23
Sample No.3
Cu conc 3.25 27.6 4.75 0.23 0.32 89.0 9.1 4.2 4.0
Pb conc 1.70 0.30 84.8 0.11 0.15 0.5 85.0 1.1 1.0
Co-Ni conc 5.38 1.17 0.91 2.70 3.85 6.2 2.9 81.2 80.4
Head (calc) 1.01 1.69 0.179 0.26
Sample No.5
Cu conc 6.84 31.2 2.32 0.25 0.32 90.9 10.5 3.2 3.2
Pb conc 1.64 0.56 78.6 0.28 0.38 0.4 85.1 0.9 0.9
Co-Ni conc 5.95 2.59 0.62 8.30 10.6 6.5 2.4 92.4 91.7
Head (calc) 2.35 1.51 0.53 0.69
50 routing at least the copper rougher tailing product
directly to the lead flotation circuit wherein a lead
concentrate is separated from a lead tailing product;
routing the lead tailing product from the lead flotation
circuit to a cobalt-nickel flotation circuit
wherein a cobalt-nickel concentrate is separated
from a cobalt-nickel tailing product; and
recovering the copper, lead and cobalt-nickel concentrates
from their respective flotation circuits.
2. The invention of claim 1, wherein the copper
rougher tailing product and first copper cleaner tailing
product are combined and routed to the lead flotation
circuit.
3. The invention of claim 1, wherein flotation of cop65
per is effected in the absence of pH modifiers other than
sulfur dioxide or sulfurous acid.
4. The invention of claim 1, wherein the primary
grind pulp is conditioned by addition of S02 in an
What is claimed is:
1. In a sequential flotation process for the separation
of components of a mineral mixture of the type wherein
a primary grind ore pulp is routed sequentially through
a series of flotation circuits having successive separation 55
and concentration stages for separating and concentrating
one of the mineral components, the improvement
comprising:
grinding a sulfide ore comprising a mixture of copper,
lead and cobalt-nickel sulfide minerals in a carbon- 60
ate matrix to provide a primary grind flotation
pulp;
conditioning the pulp with S02 under intense aeration
to depress lead and cobalt-nickel and promote copper;
routing the conditioned pulp to a copper flotation
circuit having a roughing stage and at least one
cleaning stage;
10
8. The invention of claim 1, wherein the sulfide ore is
a Missouri lead belt ore.
9. The invention of claim 1, wherein the sulfide ore is
a vibumam trend ore body of the new lead belt.
10. The invention of claim 1, wherein the sulfide ore
is located within a Mississippi Valley-type deposit.
11. The invention of claim 1, wherein the flotation of
copper is effected at an acidic pH of about 6.5 to 6.8.
12. The invention of claim 11, wherein a collector
10 highly preferential for copper in an acidic medium is
employed for copper flotation.
13. The invention ofclaim 11, wherein the collector is
ethyl isopropyl thionocarbamate.
* * * * *
4,460,459
9
amount of from about 1 to about SIbs. 802 per ton of
pulp.
S. The invention of claim 1, wherein the primary
grind pulp is intensely aerated by injection of natural air 5
into the pulp at a rate of about 3 to 5 cu ft/min.
6. The invention of claim 1, wherein lead is separated
by flotation after depression of other sulfides present
with a cyanide.
7. The invention of claim 1, wherein cobalt/nickel is
separated by flotation after activation with copper sulfate.
15
20
25
30
35
40
45
50
55
60
65
t: �Z{a;PA@sation:none;mso-layout-grid-align:none;text-autospace:none'>4,402,919
Sep. 6, 1983
• • '. •• I ,.. _ '. _ "- • •
. -
1
4,402,919
2
SUMMARY OF THE INVENTION
DESCRIPTION OF THE PREFERRED
EMBODIMENT
A feed material comprising an ore containing aluminum,
phosphorus and other values including uranium is
treated to obtain a select fraction having a particle size
less than about 150 U.S. Standard mesh and preferably
less than 200 U.S. Standard mesh to provide a concentrate
fraction. The concentrate fraction contains valuable
quantities of uranium and other elements such as
aluminum and phosphorus. The remainder of the feed
material predominately comprises quartz and is discarded.
The treatment by which the concentrate fraction
is obtained can include crushing, scrubbing, grinding
or milling of the ore to provide a particulate capable
of being sized. The particulate is sized by screening or
any other suitable means. The particular apparatus employed
can comprise any commercially available equipment
capable of producing the concentrate fraction.
fraction of large particles size quartz sand and upgrades
the unlnium, aluminum and phosphorus content of the
remainder. Acid consumption still is substantially
higher than desirable. For example, 1600 to 2000
5 pounds of 93 to 98 percent sulfuric acid are required to
dissolve 2000 pounds of concentrate. It is known that
calcining the concentrate before dissolution will reduce
acid consumption. However, acid consumption remains
at about 600 pounds of93 to 98 percent sulfuric acid per
2000 pounds of original concentrate.
It is desirable to provide a process that will permit
regeneration of a portion of the acid that is consumed to
solubilize the ore in which the uranium is present.
The surprising discovery now has been made that
uranium can be dissolved from an ore comprising aluminum,
phosphorus, uranium and other values by a procedure
which reduces the quantity of acid consumed to
effect the dissolution by over one half the quantity presently
consumed in the best prior art process. The reduction
in acid consumption is effected by regeneration of
a substantial portion of the acid consumed to solubilize
the ore.
In practice, the ore is contacted with a mineral acid to
solubilize at least a portion of the acid soluble constituents
including any uranium contained in the ore. The
ore can be physically concentrated or otherwise treated
30 such as by calcination prior to contacting the mineral
acid. The solubilization results in the formation ofa
spent acid solution containing dissolved uranium, aluminum,
phosphorus and other values together with any
undissolved solids. The spent acid solution then is
heated to a predetermined elevated temperature while
maintaining at least the autogenic pressure of the solution
to effect a precipitation of aluminum phosphate
from the solution. The precipitation results in the regeneration
of a substantial portion of the mineral acid consumed
to solubilize the ore. The uranium values then
can be recovered from the remaining solution by any
known techniques. The uranium depleted solution comprising
regenerated acid then is recycled to contact
fresh ore to solubilize additional uranium values.
Alternatively, the uranium can be recovered from the
spent acid solution prior to acid regeneration.
An additional benefit of the process is the production
of a high quality aluminum phosphate by-product.
BACKGROUND OF THE INVENTION
PROCESS FOR THE REGENERATION OF
MINERAL ACIDS USED TO SOLUBILIZE
PHOSPHATE ORES
1. Field of the Invention
This invention relates to a process for the regeneration
of mineral acids used to solubilize phosphate ores
which thereby permits recovery of uranium and other 10
valuable minerals from the ore.
2. Description of the Prior Art
It is well known in the phospheric acid technology
that phosphate ore can be treated with a mineral acid to
convert the phosphate into a soluble form, either as 15
phosphate fertilizers, phosphoric acid or phosphoric
acid compositions which can be processed into phosphate
chemicals. The solubilization process also is
known to dissolve impurities in the ore such as uranium
and vanadium which then can be separately recovered 20
from the resultant solution. One of the largest economic
expenses of the process is the cost of the mineral acid
that is consumed during the solubilization. The quantity
of mineral acid required to effect the solubilization is
directly related to the quantity of acid soluble materials 25
present in the ore. Most of the acid soluble materials are
dissolved in the process of solubilizing the phosphate
values. No simple method is known in the prior art to
regenerate the acid used to convert the phosphates into
a soluble form.
Large phosphate ore fields are known to exist in
Florida and in other areas of the United States. For
economic reasons, only the phosphate ores containing a
high ratio of phosphate to other acid soluble materials
are considered commercially recoverable. The high 35
quality commercially recoverable ores of the Florida
fields have been found to contain limited quantities of
uranium. The overburden on the high quality phosphate
ore comprises material referred to as "leached zone
material" which consists largely of sand containing 40
components of aluminum, phosphorus, iron and other
values together with clays. The leached zone material
has been formed by natural weathering or leaching of
the phosphate ore field. The low phosphate content of
this leached ore presently makes its utilization unattrac- 45
tive for the production of phosphates because of the
large quantity of mineral acid required to solubilize the
ore. However, this leached ore has been found to contain
uranium in concentrations significantly greater than
in the higher quality phosphate ore that is considered 50
commercially recoverable.
The major problem preventing the recovery of the
uranium in the Florida leached zone material and from
other phosphate ore fields is one of economics. A large
quantity of acid is required to effect dissolution of the 55
uranium present in these ores. The high acid requirement
is due to the fact that the aluminum, phosphorus
and other acid soluble values also must be dissolved to
solubilize the uranium. Further, no effective method of
physically concentrating the minerals to produce a sig- 60
nificantly higher quality concentrate for treatment has
been found.
Presently, the best known concentrating procedure
produces a concentrate of the uranium and other phosphate
minerals by scrubbing and sizing the raw ore to 65
obtain a select fraction which then is dissolved with a
mineral acid. This procedure rejects from about 60 to
about 75 percent of the ore, by weight, as a coarse
~ ~ • • "I _ -_ .. • ' .... •
. . . - . .' . ~ ..
4,402,919
3
The concentrate fraction then is admixed with a sufficient
quantity of a leach solution comprising a mineral
acid to effect solubilization of a substantial portion of
the concentrate fraction and at least a portion of the
uranium present in said concentrate fraction. The min- 5
eral acid can comprise, for example, sulfuric acid, phosphoric
acid and the like. In a typical reaction, alSO
mesh size fraction in aqueous slurry form, having a
solids content in the range of from about 30 percent to
about 60 percent, is reacted with the sulfuric acid at 10
temperatures in a range between about ambient temperature
to above the boiling temperature of the leach
solution and preferably from about 60° C. to about 90°
C. For temperatures above the boiling temperature of
the leach solution, the solubilization is effected under a 15
pressure at least equal to the autogenic pressure of the
heated solution.
Preferably, the solubilization is carried out for a period
of time ranging between 0.2 and about 15 hours and
more particularly, for a period of from about 30 minutes 20
to about 60 minutes, although the length of time may be
varied considerably depending upon other variables in
the reaction conditions. The interdependence of variables
makes for vast differences in the specific conditions
employed as to each variation. In general, it may 25
be stated the higher the percent acid acidulation used,
the shorter the time required. Thus, for example, if
about 70 percent acidulation is used, that is, about 106.5
pounds of 96 percent sulfuric acid per 100 pounds of
ore, only about 15 minutes is required to acomplish the 30
digestion, while at about 45 percent acidulation, about 6
hours digestion is necessary to give good recovery of
the desired constituents. Depending upon the analysis
of the particular ore processed, between about 30 percent
and 105 percent acidulation is desired. This corre- 35
sponds to the addition of between about 29 pounds and
about 150 pounds of sulfuric acid per hundred pounds of
ore processed. Preferably, about 70 percent acidulation
is used. The percent acidulation referred to in this description
is calculated on the basis of the reaction of 40
sulfuric acid with all of the aluminum, calcium and iron,
or other significant cationic constituents present in the
ore. In other words, 100 percent acidulation would be
the addition of that amount of sulfuric acid required to
completely react with these components. After the 45
solubilization, the aqueous solution of reaction products,
sometimes referred to as "spent acid solution," is
separated from the insolubles, such as quartz and clay.
The substantially solids-free aqueous solution of reaction
products is introduced into a reaction zone wherein 50
the solution is heated to a temperature above 100° C.
while maintaining the solution at a pressure level at least
equal to the autogenic pressure of the solution to effect
a precipitation of the leached phosphorus values as
crystalline aluminum phosphate. Preferably, the aque- 55
ous solution is heated to a temperature level in the range
offrom about 150° C. to about 200° C. and most preferably
a temperature in the range of from about 180° C. to
about 200° C. Temperatures above 200° C. can be employed
to effect the precipitation of the leached phos- 60
phorus values, however, the precipitation reaction is
essentially complete at about 200° C.
The present inventors have found that when the
leached phosphorus values are precipitated within the
aqueous solution, in the described manner, that a por- 65
tion of the mineral acid is regenerated. This is evidenced
by a significant drop in the pH level of the aqueous
solution of reaction products as the aluminum phos-
4
phate precipitate is formed. The aqueous slurry produced
at a result of the precipitation of the AIP04 also
contains other values, including uranium, that were
dissolved during solubilization of the ore. These additional
elements remain in the solution and generally do
not precipitate with the aluminum phosphate.
While the precise mechanism of the chemical reaction
involved in regeneration of the mineral acid presently
in unknown, the inventors presently believe that
the major portion of the aluminum and phosphorus
contained in the aqueous solution of reaction products,
resulting from solubilization of the ore, is in the form of
AIH2P04+2. It is believed that the mineral acid is regenerated
according to the following equation:
AIHZP04+Z+mineral acid
anion~AIP04ppl+ mineral acid
More particularly, when sulfuric acid is employed to
solubilize the ore, the acid is believed to be regenerated
according to the following equation:
An analysis of the precipitate employing x-ray diffraction
indicates that the precipitate comprises berlinite, an
anhydrous aluminum phosphate. Further, chemical
analysis of the precipitate indicates that it contains no
detectable quantity of uranium and no significant quantity
of any of the other solubilized mineral values present
in the aqueous solution of reaction products, the
precipitate being found to have a purity in excess of 99
percent aluminum phosphate. Thus, the process of this
invention also produces a high quality by-product that
has a significantly higher P20S content than, for example,
apatite, which is considered a high quality source of
phosphorus.
The precipitated aluminum phosphate can be separated
from the aqueous slurry of the same by filtration,
centrifugation gravity settling or the like. The particular
apparatus employed to effect the separation can
comprise any of that which commercially is available.
In one particular embodiment in which the mineral
acid comprises sulfuric acid, if an attempt is made to
precipitate the aluminum and phosphorus from the
aqueous solution of reaction products at a temperature
below about 100° C., a precipitate will form. However,
the precipitate is alunogen (Ab(S04h.18H20) and no
mineral acid is regenerated. When sulfuric acid comprises
the mineral acid, it also has been observed that
any calcium sulfate which may tend to precipitate from
the aqueous solution of reaction products after formation
of such solution should be permitted to form. The
precipitated calcium sulfate then should be separated
from the aqueous solution before introduction of the
now substantially solids-free aqueous solution into the
reaction zone to precipitate the aluminum phosphate.
Otherwise, a mixed calcium aluminum sulfate is found
to precipitate instead of aluminum phosphate and no
sulfuric acid is regenerated.
The presence of calcium in the aqueous solution
when mineral acids other than aulfuric acid are employed
to effect the solubilization of the ore has no
apparent effect upon the precipitation of the aluminum
and phosphorus values as aluminum phosphate. The
filtrate remaining after separation of the aluminum
phosphate, which contains dissolved uranium and other
4,402,919
EXAMPLE II
5
elements, can be treated by any known method to recover
the uranium and any other desired elements.
The uranium can be separated from the filtrate by, for
example, solvent extraction techniques whereby the
uranium values are transferred from the aqueous filtrate 5
to an organic solvent extractant. The extracted uranium
then is separated from the organic solvent by, for example,
contact with an alkaline stripping agent. Various
processes for solvent extraction of tranium and other
values from aqueous acidic solutions are disclosed in, 10
for example, U.S. Pat. Nos. 3,700,415, 3,711,591 and
3,836,476, the disclosures of which are incorporated
herein by reference. It is to be understood that the
method for separating the uranium or any other values
from the aqueous solution is not to be limited to solvent 15
extraction processes but that any method known by
individuals skilled in the art may be employed.
The practice of the process of the present invention
results in the regeneration of over 50 percent of the acid
employed to solubilize the ore. Often, the present pro- 20
cess effects regeneration of over two thirds of the mineral
acid originally employed to solubilize the ore. Such
regeneration capability permits applicants to recover
uranium present in low phosphate content ores in an
economical manner while also providing a high purity 25
by-product of aluminum phosphate which can be used
as a feed stock for production of aluminum and phosphorus
chemicals.
To further illustrate the process of the present invention,
and not by way of limitation, the following exam- 30
pIes are provided.
EXAMPLE I
A representative sample of an aqueous solution of
reaction products resulting from sulfuric acid leaching 35
of a minus 150 mesh fraction of Florida leached zone
material is introduced into a reaction zone comprising a
Parr autoclave having an acid resistant liner. The solution
is formed by contacting 1600 lbs. of 96 percent
H2S04 with one ton of uncalcined leached zone mate- 40
rial. The aqueous solution is analyzed and is found to
contain 55.3 gil Ah03, 30 gil P20S, 0.11 gil U30g and
have a pH of about 0.5. The aqueous solution is heated
in the reaction zone to a temperature of about 200· C.
while maintaining the autogenic pressure of the aqueous 45
solution. The solution· is maintained at the elevated
temperature for about 5 minutes to effect precipitation
of crystalline aluminum phosphate in the solution to
form a slurry. The slurry is withdrawn from the reaction
zone and filtered to separate the precipitate from 50
the aqueous solution. The precipitate is assayed and is
found to comprise in excess of 99 percent aluminum
phosphate and less than 0.001 percent U30g. The pH
level of the filtrate is measured and is found to be about
0.2. The filtrate is analyzed and is found to contain 23 55
gil Ah03 and 3.7 gil P20S.
The formation of the aluminum phosphate precipitate
is found to regenerate an amount of mineral acid equivalent
to in excess of 800 lbs. of 96 percent sulfuric acid
per ton of original leached zone material. This repre- 60
sents in excess of about 50 percent of the acid necessary
to solubilize a similar quantity of the leached zone material.
65
A representative sample of an aqueous solution of
reaction products resulting from sulfuric acid leaching
of a minus 150 mesh fraction of calcined leached zone
6
material is introduced into a reaction zone comprising a
modified Parr autoclave. The solution is formed by
contacting 600 lbs. of 96 percent H2S04 with one ton of
leached zone material that is calcined prior to contact
with the acid. The aqueous solution is analyzed and is
found to contain 100 gil Ah03, 120 gil P20S, 0.3 gil
U30g and have a pH of about 1.3. The aqueous solution
is heated in tlIe reaction zone to a temperature of about
200· C. while maintaining the autogenic pressure of the
aqueous solution. The solution is maintained at the elevated
temperature for about 5 minutes to effect precipitation
of crystalline aluminum phosphate in the solution
to form a slurry. The slurry is withdrawn from the
reaction zone and filtered to separate the precipitate
from the aqueous solution. The precipitate is assayed
and is found to comprise in excess of 99 percent aluminum
phosphate and less than 0.001 percent U30g. The
pH level of the filtrate is measured and is found to be
about 0.5. The filtrate is analyzed and is found to contain
39.5 gil Ah03 and 36.4 gil P20S.
The formation ofthe aluminum phosphate precipitate
is found to regenerate an amount of mineral acid equivalent
to in excess of 420 lbs. of 96 percent sulfuric acid
per ton of original leached zone material. This represents
in excess of about 70 percent of the acid necessary
to solubilize a similar quantity of the leached zone material.
While the present invention has been described with
respect to what at present are the preferred embodiments
thereof, it will be understood, of course, that
certain changes, substitutions, modifications and the like
can be made therein without departing from its true
scope as defined in the appended claims.
What is claimed is:
1. A process for the regeneration of mineral acid used
to solubilize phosphate ore comprising:
contacting an ore comprising aluminum, phosphorus
and other values including uranium with a leach
solution comprising a mineral acid selected from
the group consisting of phosphoric acid and sulfuric
acid to solubilize at least a portion thereof and
form a solution of spent mineral acid and solubilized
values in association with any non-solubilized
values, said solubilized values including aluminum,
phosphorus and uranium;
introducing said solution of spent mineral acid and
solubilized values into a reaction zone; and
heating said solution to said reaction zone to a temperature
in excess of 100· C. while maintaining the
pressure level at least equal to the autogenic pressure
of said solution to cause a substantially uranium-
free precipitate of crystalline aluminum phosphate
to form and to regenerate at least a portion of
said spent mineral acid to form regenerated leach
solution containing solubilized uranium values.
2. The process of claim 1 defined further to include
the steps of:
contacting said regenerated leach solution with an
organic extractant to extract at least a portion of
any solubilized uranium values present therein; and
recovering said extracted uranium values from said
organic extractant.
3. The process of claim 1 defined further to include
the steps of:
contacting said solution of spent mineral acid and
solubilized values, prior to heating said solution,
with an organic extractant it} I'lxtract at least l\
4,402,919
~
9. The process of claim 8 wherein the temperature to
which the solution is heated in the reaction zone is in the
range of from about 150° C. to about 200° C.
10. The process of claim 8 wherein the temperature to
which the solution is heated in the reaction zone is in the
range of from about 180° C. to about 200° C.
11. The process of claim 8 wherein at least 50 percent
of the mineral acid employed to solubilize the ore is
regenerated.
12. The process of claim 8 wherein at least 70 percent
of the mineral acid employed to solubilize the ore is
regenerated.
13. The process of claim 8 defined further to include
the steps of:
contacting said aqueous solution of reaction products,
prior to heating in said reaction zone, with an organic
extractant to extract at least a portion of any
solubilized uranium values present in said aqueous
solution; and
recovering said uranium values from said organic
extractant.
14. The process of claim 8 defined further to include
the steps of:
contacting said regenerated aqueous leach solution
with an organic extractant to extract at least a
portion of any solubilized uranium values present
therein; and
recovering said extracted uranium values from said
organic extractant.
15. The process of claim 2 defined further to include
the step of:
contacting fresh ore with the uranium-depleted regenerated
leach solution to solubilize at least a
portion of said fresh ore.
16. The process of claim 3 defined further to include
the step of:
contacting fresh ore with the regenerated leach solution
to solubilize at least a portion ofSaid fresh ore.
17. The process of claim 13 defined further to include
the step of:
contacting fresh ore with the regenerated leach solution
to solubilize at least a portion of said fresh ore.
18. The process of claim 14 defined further to include
the step of:
contacting fresh ore with the uranium-depleted regenerated
leach solution to solubilize at least a
portion of said fresh ore.
* * :(I: * *
10
7
portion of any solubilized uranium values present
in said solution; and
recovering said uranium values from said organic
extractant.
4. The process of claim 1 wherein the ore comprises 5
leached zone material.
5. The process of claim 1 wherein the temperature to
which the solution is heated in the reaction zone is in the
range of from about 100° C. to about 200° C.
6. The process of claim 1 wherein the temperature to
which the solution is heated in the reaction zone is in the
range of from about 150° C. to about 200° C.
7. The process of claim 1 wherein the temperature to
which the solution is heated in the reaction zone is in the 15
range of from about 180° C. to about 200° C.
8. A process for the regeneration of mineral acid used
to solubilize phosphate ore comprising: .
separating an ore comprising aluminum, phosphorus,
uranium and other elements into at least two frac- 20
tions, at least one of said fractions having an average
ore particle of a size capable of passing through
a U.S. Standard 150 mesh screen;
contacting said fraction capable of passage through a 25
U.S. Standard 150 mesh screen with an aqueous
leach solution comprising a mineral acid selected
from the group consisting of phosphoric acid and
sulfuric acid to solubilize at least a portion thereof
and form an aqueous solution of reaction products 30
comprising spent mineral acid and solubilized values,
said solubilized values including aluminum,
phosphorus and uranium;
separating said aqueous solution of reaction products
from any unsolubilized ore to provide a substan- 35
tially solids-free solution of reaction products;
introducing said substantially solids-free solution of
reaction products into a reaction zone; and
heating said solution in said reaction zone to a tem- 40
perature in the range of from about 100° C. to
about200° C. while maintaining the pressure level
at least equal to the autogenic pressure of said solution
to cause a precipitate of substantially uraniumfree
crystalline aluminum phosphate to fonn and to 45
cause at least a portion of said spent mineral acid to
.regenerate and fonn regenerated aqueous leach
solution containing solubilized uranium values.
50
55
60
65