4,501,721 Leaching and washing a flocculated slurry having a fiber content

Patent Number/Link:

United States Patent [19]

Sherman et at

[I I] Patent Number:

[45] Date of Patent:

4,501,721

Feb. 26, 1985

[54] LEACHING AND WASHING A

FLOCCULATED SLURRY HAVING A FIBER

CONTENT

[75] Inventors: Michael I. Sherman; Carl L. Elmore,

both of Glens Falls, N.Y.; Robert J.

Brison, Golden, Colo.

[73] Assignee: Kamyr, Inc., Glens Falls, N.Y.

[21] Appl. No.: 503,178

[22] Filed: Jun. 10, 1983

[51] Int. C1.3 COIG 7/00

[52] U.S. Cl. 423/27; 423/29;

423/109; 423/150; 75/2; 75/3; 75/101 R;

75/105; 75/Il8 R; 75/119; 75/120

[58] Field of Search 423/27, 29, 25, ISO,

423/109; 75/15 A, 3, 2, 101 R, Il8 R, 120, Il9

Refining of Leaching Residues from Zinc Electrolysis",

Erzmetall, May 1976, pp. 224-229, (English language

abstract).

Shoemaker et aI, "Recovery of Gold and Silver from

Ores", paper to International Precious Metals Institute,

10/23/80.

Siderov, "Intensification of Zinc Pulp Settling and Solution

Clarifying Through the Use of Flocculants",

Yearbook of the Institute of Non-Ferrous Metallurgy,

1978, pp. 22-37, English language abstract only.

Habashi, "Pressure Hydrometallurgy: Key to Better

and Nonpolluting Processes"; E/MJ, pp. 96-100, (2171)

and 88-94, (5/71).

Primary Examiner-John Doll

Assistant Examiner-Robert L. Stoll

Attorney, Agent, or Firm-Cushman, Darby & Cushman

FOREIGN PATENT DOCUMENTS

2085856 5/1982 United Kingdom .

OTHER PUBLICAnONS

Heinen et ai, "Enhancing Percolation Rates in Heap

Leaching of Gold-Silver Ores", International Bureau

of Mines, 1979.

Perry, "Refining Zinc Silicate Ore by Special Leaching

Technique", Chemical Engineering, 10/10/66; pp.

182-184.

Mager, "Technical and Commercial Aspects of the 14 Claims, 3 Drawing Figures

Particlized mineral material, such as gold ore, silver ore,

or coal, is subjected to a leaching process in a manner to

maximize treatment effectiveness even when the particlized

mineral material contains small fines. The material

is slurried with a flocculating material and fibers,

such as cellulosic fibers, fiberglass fibers, or ceramic

fibers, and a liquid, and then is passed to the top of a

leaching reactor. The slurry is continuously passed

downwardly in the reactor while the leaching liquid,

such as a cyanide solution, is passed counter-current to

the slurry. Leaching liquid is removed from the top of

the leaching reactor by a stilling well, and then passed

through a carbon adsorber and reintroduced into the

reactor. Leached slurry is passed to a continuous washing

station being utilized as a slurrying liquid for the

particlized mineral material.

[56] References Cited

U.S. PATENT DOCUMENTS

2,479,930 8/1946 Herkenhoff et ai. 423/29

3,151,972 10/1964 Streib 75/1

3,788,841 1/1974 Agarwal et ai. 75/103

4,071,611 1/1978 Chelson 423/41

4,174,997 11/1979 Richter 162/19

4,256,705 3/1981 Heinen et ai. 423/27

4,256,706 3/1981 Heinen et ai. 423/29

[5i] ABSTRACT

•

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u.s. Patent Feb. 26, 1985 Sheet 1 of2 4,501,721

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4,501,721

DETAILED DESCRIPTION OF THE

DRAWINGS

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 is a flow sheet illustrating the practice of a

method according to the present invention for the treatment

of gold or silver ore;

FIG. 2 is a schematic illustration of exemplary apparatus

utilizable for the practice of a method according

to the present invention; and

FIG. 3 is a detail cross-sectional schematic view of an

alternative form that a leaching reactor vessel could

take for the practice of a method according to the present

invention.

gold or silver) is continuously removed from the top of

the vessel, as by utilizing a stilling well. The leached

slurry is continuously washed, preferably in a single

stage or in a two stage wash procedure utilizing a vessel

5 or vessels comparable to the leaching reactor. For two

stage washing, the spent wash liquid from the first stage

of the washing is utilized as the liquid for slurrying the

particulized mineral material.

In order to Jock the ore particles in a stable network

for treatment in the continuous process according to the

invention, a flocculating agent and fibers are added to

the liquid and particulized mineral material during slurrying

thereof. Any suitable conventional flocculant,

such as synthetic polymers, may be utilized, and the

fibers may be selected from cellulosic fibers (e.g wood

pulp fibers), fiberglass fibers, or ceramic fibers.

The utilization of flocculants and fibers is not restricted

to the practice of a continuous process. Rather

the formation of a slurry utilizing those materials is also

applicable to batch leaching and washing operations,

and the like.

The cyanide solution leaching liquid, containing metals

removed from particlized ores, is treated in a particular

manner to facilitate removal of the metal therefrom.

The pregnant leaching solution-that removed

from the top of the leaching reactor vessel described

above-is passed through a carbon adsorbing device.

Fresh carbon is added periodically to the top of the

device and loaded carbon removed from the bottom,

and effluent from the carbon adsorption device is recycled

to the washing or leaching stage.

It is the primary object of the present invention to

provide for the effective removal of predetermined

constituents from a particlized mineral material. This

and other objects of the invention will become clear

from an inspection of the detailed description of the

invention, and from the appended claims.

A flow sheet illustrating practice of an exemplary

55 method according to the present invention is provided

as FIG. 1 of the drawings. While the invention will be

primarily described with respect to the removal of gold

or silver from gold and silver ores, it is to be understood

that the invention has broader applicability, and is applicable

to a wide variety of particlized mineral materials

to be treated with a leaching liquid. For instance, the

invention is utilizable with a wide variety of metal bearing

ores (such as copper and zinc ores), as well as for the

removal of pyritic, organic and sulfate sulfur compounds

from coal or the like.

According to the invention as illustrated in FIG. 1,

the ore is first crushed or ground and then slurried, with

flocculants and fibers added to the slurry to lock ore

BACKGROUND AND SUMMARY OF THE

INVENTION

LEACHING AND WASHING A FLOCCULATED

SLURRY HAVING A FIBER CONTENT

The leaching of constituents from particlized mineral

materials is practiced utilizing a wide variety of materials

and equipment. Leaching procedures are particularly

useful for the recovery of metals from particulized 10

mineral ores, such as gold and silver ores. The dominant

process for the extraction of such metals from ores is

leaching with alkaline cyanide solution and oxygen, and

this basic procedure has changed relatively little since

the issuance of the first patent thereon in 1887. Despite 15

extensive use of cyanide leaching, however, it has a

number of drawbacks associated therewith, including

the practical necessity of either grinding the ore fine for

continuous agitation leaching or utilizing batch leaching

methods on coarser material, significant pollution 20

loads, and minimal adaptability.

According to the present invention a method for the

leaching of particlized mineral materials to remove

constituents therefrom is provided which has enhanced

effectiveness compared to prior art procedures. The 25

method according to the invention is particularly applicable

to the removal of metals from metal bearing ores,

such as gold and silver ores, but also is adaptable to

other processes, such as the removal of the pyritic,

organic, and sulfate sulfur compounds present in a solid 30

carbonaceous fuel of the coal or coke type.

The method according to the invention effects the

treatment of a slurry of particlized mineral material in a

continuous manner, with no large pressure losses and

with good metal removal efficiency. The process is 35

capable of treating particlized mineral materials in a

continuous manner even when there is a relatively large

percentage of small fines (e.g. 200 mesh or below), without

channelling and with excellent uniformity of flow.

The process according to the invention is easily adapt- 40

able to high temperature and/or high pressure conditions,

and may be closed to the atmosphere thereby

reducing the pollution potential associated therewith.

The process also generally requires less grinding equipment

and energy and less pumping, agitating, and like 45

energy expenditures compared to prior art continuous

leaching processes, can handle ores of a wide variety of

sizes, and can handle soft materials-such coke or coal.

Further, by practicing the invention substantially the

same results can be achieved as by leaching with carbon 50

added to the "pulp" being leached, without the necessity

of adding carbon. This is in part because the retention

time for the dissolved metal (e.g. gold) in contact

with ore is very short due to countercurrent flow in the

reactor vessel of the invention.

The invention also comprises a novel slurry of particlized

mineral material, and a method of removing metals

from a pregnant leaching solution utilizing a conventional

carbon adsorber device.

According to one aspect of the present invention, a 60

particlized mineral material, such as gold or silver ore,

is slurried with a liquid. The material is passed downwardly

in a generally vertically upstanding leaching

reactor vessel, and leaching liquid is passed countercurrently

(i.e. upwardly) to the continuously down- 65

wardly passing slurry. Treated slurry is continuously

removed from a bottom portion of the vessel, while

pregnant leaching liquid (with leached constituents, e.g.

4,501,721

OPERATION

An exemplary method of operation of the apparatus

of FIG. 2, in the practice of an exemplary method according

to the present invention, will now be described.

Particlized gold ore is fed by conveyor 10 to tank 11,

to which liquid from line 13, polyall flocculant from line

14, and cellulosic fibers from line 15 are added. Sufficient

fibers are added so that they comprise between

about 0.01 percent and IO percent, by weight, of the

total slurry, preferably about 0.05-0.75%.

The slurry is continuously mixed in tank 11, and then

is continuously pumped through line 17 to enter tube 21

of the vessel 18, and continuously passed downwardly

therein. Sodium cyanide leaching liquid is introduced

through line 25 and distributing device 22 to continuously

flow counter-current to the slurry in vessel 18,

and pregnant leaching liquid is removed from the stil-

Where only a single wash vessel 27 is utilized, line 31

is connected to structure 30, and line 32 to" the tailings

pond or the like.

An alternative form the leaching reactor 18 of FIG. 2

may take is illustrated in FIG. 3, with functionally related

components in the two embodiments illustrated by

the same reference numeral, only the reference numeral

being preceded by a "I" in FIG. 3. The form of vessel

118 may, of course, be utilized for the wash vessels 27,

28, also.

For the vessel 118, the slurry feed and the treatment

liquid feed are preferably provided in concentric tubes

125, 117. The tube 117 is connected to rotating slurry

introducing device 121, which introduces slurry at level

15 120 below the overflow launder or stilling well 119

provided at the top of the vessel 118.

The liquid introduction pipe 125 is operatively connected

to the liquid distributing device 122 at the bottom

of the vessel 118, with paddles or like slurry discharge

mechanism 123 provided on the bottom of the

vessel 122. Note that in this embodiment the structures

121, 122 are rotated by a common shaft 37 connected up

to a conventional drive motor 38. A discharge control

valve 39 may be provided in the line 24 so that the

25 vessel 118 may be utilized for batch operations, as well

as for continuous operations, and to control the rate of

solids removal.

Returning to FIG. 2, a method for facilitating removal

of the metals from the pregnant leaching solution

withdrawn from the stilling well 19 will now be described.

Pregnant leaching solution is withdrawn from stilling

well 19 through line 42, and passes to the carbon adsorbing

device 43, ultimately passing to line 25.

Operatively connected to the device 43 via line 44 is

the storage tank 45, which is adapted to store strong

leaching solution therein. The invention is particularly

applicable to the leaching of gold and silver ores, in

which case the leaching liquid from source 26 comprises

a caustic cyanide solution, such as sodium cyanide,

and the solution in tank 45 is supplied from a

source 46 of hot caustic cyanide solution. Fresh carbon

is periodically added to the top of device 43 via line 47

and loaded carbon is withdrawn from the bottom

through line 44 and dumped into tank 45.

From tank 45, caustic and carbon are periodically

passed by pump 49 through valve 50 to metal removal

station 51. From metal removal station 51 carbon is

passed to carbon reactivation station 52 to supply fresh

carbon, through line 47, to device 43 as needed.

particles in a stable network. The slurry is passed to a

leach tank for counter-current flow treatment with a

leaching liquid. The pregnant solution is passed to a

station for metal recovery, while the leached slurry is

passed to wash tanks, with the residue passed to a tail- 5

ings pond, or like disposal area.

FIG. 2 schematically illustrates equipment that may

be utilizable for the practice of the novel aspects of the

process illustrated by the flow sheet of FIG. 1. The

particlized ore is passed by conveyor 10 or the like to IO

the slurrying tank 11, which may include a conventional

mixing means 12. A slurrying liquid is added to the tank

11 from line 13, the liquid in line 13 preferably comprising

water and/or the spent wash liquid from one or

more washing stages to be hereinafter described.

Also added to the slurry tank 11 is a flocculant from

source 14, and fibers from source 15. The flocculant and

fibers may be any suitable flocculant or fibers that are

capable of locking ore particles, including fines, in the

slurry in a stable network so that they may be subse- 20

quently subjected to the leaching treatment. Typical

flocculants comprise synthetic polymers of anionic,

cationic or nonionic types, and typical fibers comprise

cellulosic fibers, fiberglass fibers, ceramic fibers, and

combinations thereof.

The slurry from tank 11 is passed through line 17 to

the top of a substantially vertically elongated leaching

reactor vessel 18. The vessel may be of a type such as

disclosed in U.S. Pat. Nos. 4,061,193 or 4,174,997, having

a "stilling well" structure 19 at the top thereof 30

above the slurry introducing point 20 of a slurrry introduction

tube 21, and having a rotating liquid introducing

device 22 at the bottom thereof, preferably commonly

rotatable with slurry discharge structure 23.

The slurry flows continuously downwardly in vessel 35

18, and is ultimately discharged through line 24 at a

bottom portion of the vessel 18. Leaching liquid is introduced

through line 25-as from source 26-so that it

flows to the distributor 22, and then flows upwardly in

vessel 18-counter-currently to the slurry flow therein. 40

The fibers and flocculant in the ore slurry lock the ore

particles in a stable network so that leaching can occur

without channelling, and with a minimal pressure drop

over the vessel 18. Thus fines of 200 mesh, or smaller,

can be handled without substantial difficulty. Addition- 45

ally, ores having particle sizes up to i inch mesh, or

even larger, may be handled at the same time that the

fines are being handled.

The leached slurry withdrawn in line 24 passes to a

washing station. At the washing station, preferably a 50

single washer 27 is provided, although a first washer

(27) and a second washer 28 (or more) may be provided.

The vessels 27, 28 are substantially identical to the vessel

18, but generally of smaller size, and include a stilling

well arrangement 29, 29' at the top thereof, and a 55

rotating liquid introducing structure 30, 30' at the bottom

thereof. Where two vessels are utilized, clean wash

water enters vessel 28 through line 31, passes through

liquid distributor 30' counter-current to the slurry flow

in vessel 28, and spent wash liquid is withdrawn from 60

the stilling well 29' through line 32, to be used as feed

wash liquid for the first wash vessel 27. Spent wash

liquid in vessel 27 withdrawn from stilling well 29

through conduit 33 preferably is passed to the slurrying

tank 11, providing a liquid feed to line 13. The washed 65

slurry is withdrawn from second wash tank 28 through

line 34, and is passed to a tailings pond, or like disposal

site.

4,501,721

ling well 19-without the necessity for utilizing screens-

in line 42.

The pregnant solution in line 42 passes through the

carbon adsorber device 43, and then is recirculated

through line 25 to the device 22. Fresh carbon is sup- 5

plied to, and loaded carbon removed from, device 43 as

necessary.

Leached slurry is discharged by device 23 from the

bottom of the vessel 18 into line 24, and passes to the top

of washing stage 27. The slurry flows downwardly 10

through washing stage 27, and then may be fed to the

top of vessel 28, and ultimately passes through discharge

34 to a disposal site. Wash water introduced in

line 31 to vessel 28 flows counter-currently to the

slurry, and the spent wash liquid withdrawn from line 15

33 is pumped to the slurry in tank 11.

Utilizing the equipment illustrated in FIGS. 2 and 3,

it is easy to adapt the process to high temperature and

high pressures, thus providing versatility in the processes

which may be practiced, and in the available 20

leaching liquids. Also, the process can be closed to the

atmosphere thereby minimizing discharges of pollutants

into the atmosphere. Because little agitation of the

slurry is necessary (only that small amount provided in 25

tank 11), energy requirements are minimized, and the

process is readily adaptable to the handling of "soft"

particlized mineral materials, such as coal. Further, the

process can be operated in a continuous manner even

when a relatively high percentage of fines, including 30

small fines, are present, and may be practiced with particles

up to about ~ inch in diameter.

While the invention has been herein shown and described

in what is presently conceived to be the most

practical and preferred embodiment thereof, it will be 35

apparent to those of ordinary skill in the art that many

modifications may be made thereof within the scope of

the invention, which scope is to be accorded the broadest

interpretation of the appended claims so as to encompass

all equivalent methods and products. 40

What is claimed is:

1. A method of removing metal from a particlized

metal bearing ore, utilizing a treatment vessel, comprising

the steps of:

(a) mixing the particlized ore with a liquid to form a 45

liquid slurry, a flocculating material so as to lock

the particlized ore in a stable network in the slurry,

and fibers so as to facilitate locking of the particlized

ore in a stable network in the slurry;

(b) continuously passing the slurry downwardly in 50

the vessel;

leaching the metal to be removed from the metal

bearing ore, upwardly in the vessel, countercurrent

to the slurry passage, to remove metal from the ore 55

in the slurry;

(d) continuously removing the treatment slurry from

a bottom portion of the vessel;

(e) continuously removing pregnant leaching liquid,

with metal removed from the ore, from a top portion

of the vessel; and

(f) continuously washing the slurry removed in step

(d).

2. A method as recited in claim 1 wherein the flocculent

is selected from the group consisting essentially of

synthetic polymers of anionic, cationic, and nonionic

types.

3. A method as recited in claim 1 wherein the fibers

are selected from the group consisting essentially of

cellulosic fibers, fiberglass fibers, ceramic fibers, and

mixtures thereof.

4. A method as recited in claim 3 wherein the flocculent

is selected from the group consisting essentially of

synthetic polymers of anionic, cationic, and nonionic

types.

5. A method as recited in claim 4 wherein the fibers

comprise, by weight, between about 0.01 percent and 10

percent of the slurry.

6. A method as recited in claim 5 wherein the fibers

comprise, by weight, between about 0.05 and 0.75 percent

of the slurry.

7. A method as recited in claim 4 wherein the leaching

liquid in step (c) is a cyanide solution.

8. A method as recited in claim 1 wherein the fibers

comprise, by weight, between about 0.01 percent and 10

percent of the slurry.

9. A method as recited in claim 8 wherein the fibers

comprise, by weight, between about 0.05 and 0.75 percent

of the slurry.

10. A method as recited in claim 1 wherein step (f) is

practiced utilizing a washing vessel, and by passing

wash liquid upwardly in the vessel counter-current to

slurry moving downwardly in the vessel, with spent

wash liquid removed from a top portion of the vessel

and washed slurry removed from a bottom portion of

the vessel.

11. A method as recited in claim 1 consisting essentially

of said steps (a)-(f), so that carbon addition to the

slurry is not practiced.

12. A method as recited in claim 10 wherein said

washing vessel comprises a first vessel, and wherein

step (f) is practiced utilizing a second vessel substantially

identical to the first vessel, the slurry outlet from

the first vessel being connected to the slurry inlet to the

second vessel, and the wash liquid inlet to the first vessel

being connected to the wash liquid outlet from the

second vessel; and wherein the wash liquid outlet from

the first vessel is used as a liquid for practicing step (a).

13. A method as recited in claim 1 wherein the leaching

liquid in step (c) is a cyanide solution.

14. A method as recited in claim 13 comprising the

further step of: (g) continuously passing pregnant liquid

withdrawn in step (e) through a carbon adsorbing device,

and reintroducing the liquid into the vessel as

leaching liquid for step (c).

* * * * *

imes zw�^RmPA@sf";mso-fareast-font-family: HiddenHorzOCR'>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.