4,460,459 Sequential flotation of sulfide ores

Patent Number/Link:

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

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

CONCENTRATE

4,460,459

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

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.

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

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

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

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

Primary grind

5" X 12" batch mill

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

Roughers

NaCN

0.05

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

4,460,459

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

0.2

0.1

0.05

0.04

0.8 0.3

Cu regrind

5" X 7" pebble mill

0.016

0,02 0.015

Roughers

.1000g 0·1

Cleaners

250 g 0·1

6:5

6.7

2 6.3

8.5

Ca(OHh

0.05

0.Q2

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

1 4 8.0

2 4

1 4 8.0

1 3

1 2

0.6

0.2

CUS04 MIBc4

0.05

0.01

Co, Ni 1st cleaner

500g 0·1

1300

0.05 0.01

0.025

Reagents Added, PoundslTon

0.05

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

4,460,459

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

Rougher

lOOOgD-1

1800

NaCN

0.05

0.025

0.05

0.025

0.D25

0.01

0.5

0.05

0.05 0.2

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;

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

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.

t: �Z{a;PA@sation:none;mso-layout-grid-align:none;text-autospace:none'>4,402,919

Sep. 6, 1983

• • '. •• I ,.. _ '. _ "- • •

. -

4,402,919

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

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-

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

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.

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

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: * *

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.