4,556,422 Process for the recovery of lead and silver chlorides

Patent Number/Link:

United States Patent [19]

Reynolds et al.

[11] Patent Number:

[45] Date of Patent:

4,556,422

* Dec. 3, 1985

[54] PROCESS FOR THE RECOVERY OF LEAD

AND SILVER CHLORIDES

[75] Inventors: James E. Reynolds, Golden; Alan R.

Williams, Denver, both of Colo.

[73] Assignee: Hazen Research, Inc., Golden, Colo.

[ * ] Notice: The portion of the term of this patent

subsequent to Jun. 30, 1998 has been

disclaimed.

[21] Appl. No.: 255,649

[22] Filed: Apr. 20, 1981

1,951,342 3/1934 Bradley 75/77

3,172,753 3/1965 Walsh 75/77

3,764,490 10/1973 Chambers 75/101 R

4,082,629 4/1978 Milner , 423/98

4,135,993 1/1979 Um 423/98

4,276,084 6/1981 Reynolds 423/98

OTHER PUBLICATIONS

International Critical Tables, vol. 7, (1st ed.),

McGraw-Hill, N.Y., 1930, pp. 313-316.

Primary Examiner-John Doll

Assistant Examiner-Robert L. Stoll

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

27 Claims, No Drawings

A process for selectively leaching lead and silver chlorides

from a sulfide ore residue in a rapid time which

comprises brine leaching the residue under pressure at a

temperature above the normal boiling point of the solution,

preferably above 100° C.

Modifications are leaching at the agglomeration temperature

of sulfur when present in the residue to agglomerate

the sulfur for ease of recovery, and flashing

from leach temperature to ambient as a lead chloride

crystallization recovery step to produce a large crop of

lead chloride crystals per pass.

Related U.S. Application Data

[63] Continuation-in-part of Ser. No. 80,444, Oct. I, 1979,

Pat. No. 4,276,084.

[51] Int. Cl.4 C22B 13/04; COlG 21116;

COlG 5/02

[52] U.S. Cl 75/101 R; 75/114;

75/120; 75/118 R; 75/78; 423/27; 423/39;

423/94; 423/98

[58] Field of Search 423/27, 39, 98, 94;

75/118 R, 120, 77, 78, 101 R, 114

[56] References Cited

U.S. PATENT DOCUMENTS

1,396,740 11/1921 Ganelin , 423/98

[57] ABSTRACT

4,556,422

PROCESS FOR THE RECOVERY OF LEAD AND

SILVER CHLORIDES

RELATED APPLICATIONS

This application is a continuation-in-part of my U.S.

patent application flIed Oct. 1, 1979, Ser. No. 080,444,

U.S. Pat. No. 4,276,084 entitled "Hydrometallurgical

Process to Recover Lead from Lead Concentrates."

TECHNICAL FIELD

The invention lies in the field of recovery of the

chlorides of lead and silver by selectively solubilizing

the chlorides from other solid materials and the final

recovery of the metals from the solubilized chlorides.

BACKGROUND ART

silver chlorides, it has been found effective for selectively

leaching lead and silver chlorides at a rapid rate

from residues obtained by cupric chloride leaching of

pyritic or sulfidic lead ores in which lead and silver

5 chlorides are selectively leached at saturation into solid

products, all as disclosed in my continuation-in-part

application referred to above. Among other materials,

these residues contain the sulfides of copper, iron and

zinc. The process is equally effective for selectively

10 leaching the residues produced from halogen leach of

pyritic ores in the processes of U.S. Pat. Nos. 4,135,993

and 4,173,623. The operation of the invention will be

illustrated by its application for the recovery of lead

from the residue resulting from the cupric chloride

15 leach of sulfidic ores described in the above-referred-to

continuation-in part application.

Pressure Brine Leach

The solids from the cupric chloride leach reaction,

comprising lead chloride, silver chloride, elemental

sulfur, unreacted metal sulfides of other metals, and

gangue, are treated for the selective separation of lead

and silver chlorides. The lead and silver chlorides are

selectively solubilized from the residue in the illustrative

embodiment by leaching with an aqueous brine

solution having a sodium chloride concentration of

from about 200 gpl to saturation. Suitable substitues for

sodium chloride are the other alkali metal chlorides,

lithium and potassium chlorides, as well as the alkaline

earth metal chlorides, calcium and magnesium chlorides.

When other solutes than sodium chloride are used

the upper limit of the amount used will change, the the

minimum amount of solute preferably being above

about 200 gpl. A leach pH of about 0-7 is preferred. Use

of too high a pH will precipitate lead compounds. Hydrochloric

acid (hydrogen chloride) may also be used as

one of the chlorides. The solute must be a chloride

which provides maximum chloride ion concentration to

40 the saturation point under the reaction conditions. The

solubilization of silver chloride may be enhanced by the

use of an oxidant in the leach, such as, sodium chlorate

or oxygen.

The brine leach is conducted at a temperature in

excess of the solution boiling temperature, which, of

course, requires a pressurized system. The temperature

is maintained between about 100· C. to about 170· C.,

the system pressure being selected so as to accommodate

the solution temperature while preventing solution

boiling. Pressures from about 30 to about 150 psig are

suitable to accomplish this purpose. If elmental sulfur is

present in the residue the agglomeration temperature of

sulfur is used. This is about 130· C. to 140· C. It was

found that when the sulfur is agglomerated during the

leach and separated from the liquid chlorides in this

form with the other solids it can be readily separated

from the other solids by physical methods, such as, wet

screening.

The brine leach, conducted under the described temperatures

and pressures, accomplishes a relatively high

solubility of lead and silver chlorides in a relatively

short period of time, while leaving the elemental sulfur

and unreacted metal sulfides in the residue phase. Retention

times of from about 30 seconds to about 5 min-

65 utes are generally adequate to dissolve lead chloride to

solution concentrations of at least about 130 grams per

liter of lead. A preferred leach time is not in excess of

about three minutes. Increased lead concentrations as a

DISCLOSURE OF THE INVENTION

Prior Art Statement

U.S. Pat. No. 4,113,471 discloses a process for brine 20

leaching oxide ores to non-selectively solubilize nonferrous

metal values as chlorides, at elevated temperatures

and pressures with the addition of oxygen. Leaching

is required for a time of ! to 12 hours. There is no

:~~~~~v~r~~~ching of lead or silver chlorides from their 25

U.S. Pat. Nos. 4,135,993 and 4,173,623 teach brine

leaching lead chloride at temperatures of 80·-120· C. to

selectively solubilize the lead chloride out of a sulfidic

residue containing the sulfides of copper, iron and zinc

as well as elemental sulfur. The leaching is not done 30

under pressure. Leach times of ! to 2 hours are required

in both processes.

International publication No. WO 80/00852, class 22

B 13/00, published under the patent cooperation treaty

on May 1, 1980 (01.05.80), discloses the recovery oflead 35

from crystallized lead chloride by reduction with hydrogen

or a hydrogen containing compound accompanied

by condensing lead chloride volatilized during the

reduction process.

BEST MODE FOR CARRYING OUT THE

INVENTION

Although the process is not limited in its application

to any particular starting materials containing lead and

Lead and silver chlorides are selectively separated

from other solid materials by leaching the materials

with a brine leach at elevated temperatures and pressures

to selectively solubilize the chlorides, followed by 45

a liquid-solids separation. A starting material on which

the process is particularly effective is the sulfidic resi- .

due obtained by selectively leaching a sulfidic lead ore

with a cupric or a halogenating leach to produce solid

lead chloride. Temperatures ranging from the boiling 50

point of the solution to 170· C. are used to selectively

solubilize substantially all of the lead and silver chlorides

in a rapid time which can be not in excess of about

three minutes. If elemental sulfur is present in the starting

material, it is agglomerated at elevated temperatures 55

for ease in separation from the solid residue. Lead chloride

is crystallized from solution by flashing from the

high temperature of the leach to lower temperatures to

produce a large crop of lead chloride crystals per pass

and elemental lead of high purity recovered from the 60

crystallized lead chloride by hydrogen reduction, or

otherwise. Silver chloride is recovered from the mother

liquor and processed for recovery of silver.

4,556,422

Wt%Pb Extraction (%)

26.1 gil

20.6%

28.6 gil

18.7%

28.5 gil

17.2%

Vol.

120ml

1.9 g

IlOml

I.3g

625 ml

9.1 g

TABLE 2-continued

Residue

Distribution

Size Weight S' Weight S'

Fraction g % % %

Plus 200 (beads) 1.67 81.8 33.9 89.5

Minus 200 (fines) 3.26 ~ ~ ~

Totalloverall 4.93 (31.0) 100.0 100.0

Rate of Brine Leaching Test

galena in L14 liter of 90 gil Cu+2 and 200 gil

NaCI at pH = I (HCI) for one hour

at 60' C.

Brine Leach Solution: I liter 250 gil NaCl, pH 1.5.

Procedure: Feed added to solution with continuous

stirring. Thief samples removed at

designated time and immediately vacuum filtereg

without rinsing. .

Leach

Time

(min.)

40 5.0 g, S' - agglomerated product from Test 1151-107-1:

Conditions: Test 1151-99-1 leach residue

130' C.

pH 11.8 with KOH

2 Hours

Elemental Sulfur Distribution in ± 2oo-mesh Size

Fractions of S'-Agglomerated Autoclave Leach Residue

Feed to Wet Screening

Tem- Sample Diluted Diluted Total Pb

per- Vol- to Sample Pbin Solubilature

ume Volume Pb Sample ity

'C. ml ml gil g gil

42 10.0 200 0.746 0.149 14.9

70 10.0 200 1.45 0.290 29.0

94 10.0 200 2.94 0.588 58.8

1001 87.5 795 9.41 7.48 85.5

1241 76.0 1360 6.20 8.43 110.9

1441 79.0 1945 7.13 13.87 175.5

47.5 10.0 200 0.972 0.194 19.4

69.0 10.0 200 2.97 0.594 59.4

96 10.0 200 6.02 1.204 120.4

1001 8.5 1320 8.96 11.8 139.1

1241 8.3 2060 6.67 13.7 165.5

1352 9.2 1625 11.0 17.9 194.3

1431 8.2 2200 8.82 19.4 236.6

240

320

NaCI

gil

result of high temperature and pressure brine leach

significantly facilitate further separation. processing.

The brine leach may be conducted at lower pressures,

including atmospheric pressure, and lower temperatures,

as in the prior art. However, pressures and tern- 5

peratures lower than those recited for the preferred

range of the brine leach will require more of the brine

solution per amount of lead and silver chlorides and a

longer retention time in order to solubilize the chlorides.

In a·typical application, washed tails or residue as. a

filter cake. from the above-referred-to cupric leach of

sulfidic ores was brine leached in an externally heated

concentric double pipe pressure leach. Brine containing

280 gil of NaCl lmd PbC1zcake was heated to 135° C. 15

under 50 psig pressure with a one minute retention time

to dissolve PbC12UP to a concentration of 145 gil Pb.

This procedure reduces the size of the crystallizer used

in subsequent.PbCh crystallization and circuit •flows to

a fraction of that of an ambient pressure system with a 20 Sulfur Agglomeration

corresponding reduction in heating and cooling needs. The agglomeration of sulfur is accomplished during

To explore the •effectiveness .of. the brine pressure leaching. by operating the brine leach within the sulfur

leach at high temperatures in rapidly solubilizinglarge agglomeration temperature range thereby permitting

amounts of lead chloride, the solubility system PbCI2- the sulfur to be readily separated from the remainder of

NaCI-H20 was extended to 144°C. at two hrinecon· 25 the residue following liquid-solids separation. Agglomcentrations

of 240 and 320 gil of NaCI used for leaching eration tests were run on a brine leach residue from

a lead sulfide ore residue from a cupric leach, the com- cupric. leach of lead sulfide ore as referred to above

position ofthe residue being typified by the brine leach containing elemental sulfur to see if the sulfur could be

residues of Examples 1.and 2. The results of the solubil- coalesced to a size large enough for a wet screen separaity

tests reported in Table 1 below show a decided 30 tion. The autoclave leach was made at 130° C. The

nonlinear increase in the solubility oflead chloride with results recorded. in Table 3 below shows that the plus

increase in temperature above the. boiling point. of the 200 mesh fraction contains about 90 percent•ofthe free

solution,and particularly above 125°C. sulfur with a grade of 82 percent, thus shoWing that the

TABLE 1 35 procedure is feasible for sulfur separation.

------,-:.;:...:::.::.::...:-------- TABLE 3

'Sample withdrawn from pressurized autoclave (150 psig N2) using sample bomb. 50

2Repeat run to check data, new solutions.

Since a pipeline brine leach is contemplated in the

most feasible commercial application of the process, a

minimum leach time is required in the interest of reducing

equipment cost and processing time. Rate of brine

leaching tests at high temperatures were made on a

residue obtained by the above-referred-to cupric leach

of a lead sulfide ore. The brine leach contained 250 gil

of NaC1 and a pH of about 1.5 was used. The leach

temperature was 140° C. The results recorded in Table

2 below indicate that substantially all of the PbClz is

leached in a time not in excess of about three minutes.

Liquids-Solids Separation

Following the brine leach, the pregnant lead and

silver chloride solution is separated from the remaining

55 residue for subsequent recovery therefrom of lead and

silver values. As high temperatures and pressures are

utilized during the leach, the liquid-solid separation

must be conducted under pressure in order to prevent

flash crystallization of the lead chloride from the solu-

60 tion. One suitable technique to accomplish the separation

while avoiding flash crystallization is to employ

small diameter pressurized liquid cyclones in parallel,

the hydroclones operating to permit pressure reduction

TABLE 2 to atmospheric as the cyclone operation effects a liquid-

--------....::~::..::=--=---------65 solids separation. A pressure drop of about 40 psi across

Rate of Brine Leaching Test the cyclone system occurs. Hydroclone techniques such

Feed: 50 g cu+

grade galena contlaeiancihnegd5r8e.6si8dupeerocfenathPigbhand· as those discussed in The Hydroclone, D. Bradley, Perobtained

by leaching 200 g of a high grade gamon Press, Lrd. 1965 may be utilized in this context.

4,556,422

material which is substantially impervious to the corrosive

action of lead chloride, such as, castable or refractory

brick. The materials of which the reactor walls

must be made have such a low heat conductivity that it

is practically impossible to heat the reactor contents

with heat applied to the outside of the walls. Accordingly,

it was necessary to devise a practical procedure

for internally heating the reactor contents to a temperature

up to 900· C. at least. Two alternative procedures

were found to be feasible.

In accordance with one procedure a furnace or reactor

made of refractory brick was used. Heat for the

endothermic reaction occurring in the reaction chamber

was supplied by fire tubes submerged in molten lead

in contact with lead chloride and the other reactants in

the reaction chamber. Means are provided for introducing

reactants into the reaction chamber and for continuously

or intermittently tapping pure lead from the furnace.

Means are also provided for condensing vaporized

lead chloride and returning the vaporized lead

chloride to the reaction chamber. Lead chloride does

not react with molten lead and having a lesser specific

gravity floats on top of the molten lead.

The second procedure comprises introducing into the

reaction chamber a partially uncombusted gas mixture

supplying hydrogen, and completing the combustion

with oxygen in an endothermic reaction which supplies

heat for the endothermic lead chloride reduction reaction.

Heat balance calculations showed that sufficient

30 heat can be brought into the lead chloride reduction

reactor to supply the endothermic heat of reaction and

other heat requirements, including that caused by heat

loss, by using a reducing combustion gas or gas mixture,

the term "gas" as used herein and in the claims including

both. Any hydrocarbon or mixture of hydrocarbons

which supply hydrogen can be used. A mixture produced

by a partial combustion ofhydrocarbons, such as,

methane or propane, provides both the hydrogen and

the heat needed for the endothermic reduction of lead

chloride. Contrary to what might be expected, introduction

into the reaction area of large volumes of water

vapor formed in the partial combustion reaction and

diluent gases does not adversely affect the reduction

reaction. The above described procedure applies also to

the recovery ofcopper from cuprous chloride by reduction

of hydrogen.

Illustrative gases and gas mixtures found suitable are

H2-CO-C02, H2-CO-C02-H20, and H2-CO-

N2. The gases used mayor may not be supplemented

by hot reducing combustion gas. Oxygen gas or

air may be used to supply oxygen.

The PbCh cake was metered to a brick-lined PbCh

reduction furnace as described above operating between

600·-900· C., preferably at about 800· C. A reducing

gas feed of 98 percent H2 from an on-site H2

plant was used. An excess of 240 percent of theoretical

H2 was fed based on lab tests in batch tube furnace runs.

This produces an exit gas consisting of 60 percent HCI

and 40 percent H2, by volume. Some volatilized PbCh

leaves the reactor zone with the off-gas but is refluxed

back to the furnace by either a molten lead splash condenser

or an air-cooled surface condenser. Any additional

heat requirements for the endothermic reduction

reaction and to bring reactants up to temperature may

be supplied by indirect firing of submerged fire tubes in

the molten lead in the reactor as described above. High

purity lead is tapped continuously or intermittently

from the furnace into a casting machine.

Another solids separation device, such as, an insulated

or jacketed pressure leaf filter can be used to accomplish

the same objective.

In operation, the pipeline dissolver discharges

through a bank of 10 mm alumina cyclones to remove 5

solids at about a 4-5 micron cut point with a let down

from a 50 psig pipeline leach to atmospheric, the pressure

being utilized to remove the solids. Flocculant may

be injected at the cyclone inlet to improve clarity of the

cyclone vortex flow. The apex flow, containing unre- 10

acted sulfides and agglomerated sulfur, flashes to atmosheric

pressure and mixes with concentrate and

mother liquor from the subsequent PbCh crystallization

to quench the hot slurry and solidify beads of agglomerated

sulfur. The slurry is gravity-fed to a wet screen or 15

similar separation device to make a separation of agglomerated

sulfur beads from other solids, principally,

unreacted sulfide tails. The fines are dewatered and

fmally filtered by conventional filtration. Filtrate is

recycled to the leach feed tank and tails cake is dis- 20

charged to a solids disposal area. Prior to reaching the

leaching tank the leach can be purified by a bleed

stream in which copper and lead values are recovered

by iron cementation and soda ash used at pH 9 to precipitate

Fe, Mg, and Zn to permit recycle of barren 25

brine. The residue from sulfur separation is disposed of

or further processed for recovery of metal values if

warranted.

Lead Chloride Crystallization

Lead chloride is crystallized from the liquid phase

resulting from the liquid-solids separation on the brine

leach solution for subsequent recovery of elemental

lead by hydrogen reduction, or otherwise. Two-stage

crystallization may be used with the first-stage at atmo- 35

spheric pressure and the second stage at about 50 mm

Hg absolute to cool the feed to about 40· C. A pregnant

brine containing up to 145 gil Pb flashes typically from

135· C. to ambient temperature in the first stage to

produce a large crop of crystals per pass. Surface con- 40

densers may be used for the second stage, with contaminated

lead chloride condensate being recycled to process.

Mother liquor overflow and crystal withdrawal

elution leg are specific design requirements to elute

minus 5-micron impurities not removed in the cyclones. 45

Alternatively, polish fIltration techniques could be used

to separate minus 5-micron solids.

Crystallizer under-flow is removed through an elution

leg at 40-50 percent solids and advanced to a washing

centrifuge. A three percent moisture PbCh cake is 50

conveyed to a surge hopper above the PbCh reduction

furnace.

Lead Recovery from Lead Chloride

The lead chloride is reduced to high purity lead by 55

hydrogen directly without further refining. The remaining

solubilized silver chloride is treated for recovery of

silver by cementation or other means. Other conventional

methods may be used to recover elemental lead

from the lead chloride. Hydrogen supplying com- 60

pounds, such as, methane and propane may be used as a

source for hydrogen.

Since the reduction of PbCh is endothermic, heat

must be supplied to the reaction represented by the

formula PbCh+H2Pb+2 HCI when an excess of hy- 65

drogen over stoichiometric is used. As lead chloride is

extremely corrosive, the reactor cannot be made of

conventional reactor materials but must be made of

2.0 0.33

Zn Fe

Brine Leach Residue Assay, %

Cu (oz/ton)

We claim:

1. A process for solubilizing a chloride selected from

the group consisting of lead chloride and a combination

of silver and lead chlorides which comprises subjecting

the chloride to a brine leach at a temperature above the

normal boiling point under a pressure above atmospheric

sufficient to prevent boiling for a sufficient time

to solubilize substantially all of the chloride.

2. The process of claim 1 in which the temperature is

maintained above the normal boiling point to about 1700

3. The process of claim 1 in which the time of leach

sufficient hydrochloric acid to maintain a pH of about 1.

The cupric chloride leach was conducted at a temperature

of 600 C. for two hours. The residue of the cupric

chloride leach was subjected to a 900 milliliter brine

5 leach at a temperature of 800 _900 C. and about atmospheric

pressure of one-half hour. The brine solution

contained about 250 grams of sodium chloride per liter.

The analysis of the brine leach residue, which weighed

83 grams, and the results of the extraction are set forth

10 in Table 5.

The extraction resulted in 19.0 grams oflead chloride

being produced. This lead chloride was reduced to lead

in an atmosphere of 175 cubic centimeters per minute of

hydrogen, 75 cubic centimeters per minute of carbon

15 monoxide, 75 cubic centimeters per minute of carbon

dioxide at a temperature of 8000 C. for 35 minutes. The

lead metal was assayed by emission spectroscopy. The

lead metal was 99.98 percent pure. It contained impurities

of 0.01 percent silicon, 0.005 percent iron, 0.001

percent copper and 0.001 percent bismuth with no other

elements being detected.

TABLE 5

4,556,422

EXAMPLE I

Off-gas is scrubbed in a packed tower or similar

scrubbing device using liquor from cupric leach, and a

large excess of dilution air to lower H2 content to a safe

level and also simultaneously consume scrubbed HCl

which may be used to reoxidize the cuprous ion to

cupric. A water scrubber may also be used to recover

the HCI. Exit gas, free of HCI and particulate matter, is

exhausted to atmosphere.

Up to over 99 percent of lead was obtained from the

starting material. Lead having a purity of +99.9 percent

was consistently obtained by the process. The

recoveries of lead and silver shown in Tables 4 and 5, as

produced by Examples 1 and 2, are representative of

recoveries obtained by the process. The lead purity

obtained in Example 2 is also typical.

Two different 100 gram samples of a lead concentrate

having a composition of 18 percent lead, 26.2 percent

zinc, 0.54 percent copper, 5.1 troy ounces of silver ore 20

ton of concentrate, 0.029 percent antimony and 14.4

percent iron were treated with 250 milliliters of a cupric

chloride leach solution comprising about 50 grams of

copper per liter as cupric chloride and 200 grams of

sodium chloride per liter. The pH of the leach solution 25 Pb

was maintained at about 1 through the addition of hydrochloric

acid. After 3 hours, a total of 4.08 and 4.80 0.21 0.805 Ex~:~~tion. 2~8

grams ofhydrochloric acid were added to Sample 1 and Pb Cu Zn Fe Ag Sb .

Sample 2, respectively. The cupric chloride leach of

Sample 1 was conducted at a temperature of 600 C. and 30 __9_9_.4 -_4..,;8;,...5 3._6__7_._1 8_1._6 ;.;32;...-_

the cupric chloride leach of Sample 2 was conducted at

a temperature of 800 C. The residue of the cupric chloride

leach of each of the samples was separately brine

leached in a brine solution containing about 250 grams

of sodium chloride per liter at a temperature of 800 _850 35

C. and about one atmosphere for one-half hour. Each

brine leach slurry was filtered while hot and the residue

was washed fIrst with hot brine solution and then with

water. The analyses of the brine leach residue and the

results of this extraction are set forth in Table 4. The 40

negative extracted copper percentages are due to a

portion of the cupric chloride of the leach solution

being precipitated to copper sulfIde.

TABLE 4

Cupric

Leach Brine Leach Residue Assay, %

Time Weight Ag Extraction, %

(Hours) Product (gm) Pb ZN Cu FE (oz/ton) Sb Pb Zn Cu Fe Ag Sb

Sample 1:

1.0 1hr. 4.32 0.49 36.1 0.78 20.3 1.4 0.013 98.3 11.8 7.9 7.6 82.4 71

residue

3.0 Final 50.7 0.11 36.1 0.96 20.0 1.8 0.014 99.6 18.7 -4.5 18.1 79.2 72

residue

Sample 2:

1.0 1hr. 4.43 0.10 37.4 0.98 19.8 1.2 0.Q18 99.6 7.2 -17.5 10.6 84.7 60

residue

2.0 2 hr. 4.52 0.08 36.7 1.00 19.2 1.1 0.Q11 99.7 17.4 -8.9 21.3 87.3 78

residue

3.0 Final 48.3 0.06 36.2 1.44 19.8 1.2 0.Q15 99.8 22.6 -49 23.0 86.8 71

residue

EXAMPLE 2

A 125 gram sample of a lead concentrate having a

composition of 25.1 percent lead, 9.57 percent zinc, 0.36

percent copper and 16.3 percent iron was treated with

500 milliliters of a cupric chloride leach solution comprising

about 50 grams of copper per liter as cupric

chloride, 200 grams of sodium chloride per liter and

is not in excess of about three minutes and a concentration

of at least about 130 gpl oflead chloride is obtained.

4. The process of claim 1 in which the brine leach

65 comprises an aqueous solution containing about

200-300 gpl to saturation of a soluble chloride which

provides a maximum concentration of chloride ion

below saturation.

4,556,422

5. The process of claim 4 in which the chloride is a

member selected from the group consisting of chlorides

of alkali and alkaline earth metals and hydrogen chloride.

6. The process of claim 1 in which the chloride solu- 5

bilized is lead chloride.

7. A process for selectively solubilizing a chloride

selected from the group consisting of lead cWoride and

a combination of lead and silver chlorides contained in

a mixture of other solids including metal sulfides which 10

comprises subjecting the mixture to a brine leach under

a normal boiling point.

8. The process of claim 7 in which the temperature is

in excess of about 100· C.

9. The process of claim 8 in which the leaching is 15

performed in a time not in excess of about three minutes

and a concentration of at least about 130 gpl of lead

chloride is obtained.

10. The process of claim 8 in which said metal sulfides 20

include the sulfides of copper, iron and zinc.

11. The process of claim 10 in which the time of leach

is less than about three minutes and a concentration

tration of at least about 130 gpl of lead chloride is obtained.

12. The process of claim 10 in which the brine leach

comprises an aqueous solution containing at least about

200 gpl of soluble chloride which provides a maximum

concentration of chloride ion below saturation.

13. The process of claim 12 in which said cWoride is 30

a member selected from the group consisting of chlorides

of alkali and alkaline earth metal cWorides and

hydrogen cWoride.

14. The process of claim 8 in which the chloride

solubilized is lead cWoride. 35

15. The process of claim 13 in which said chloride is

sodium cWoride.

16. The process of claim 8 in which the solubilized

silver and lead cWorides are separated from solids.

17. The process of claim 16 in which said separation 40

is accomplished by liquid cyclone separation.

18. :The process of claim 16 in which lead chloride is

crystallized from the solution and silver chloride is

recovered from the mother liquor.

19. The process of claim 16 in which lead chloride is

recovered from the brine solution by crystallization.

20. The process of claim 19 in which said crystallization

includes flashing from the brine leach temperature

to a lower temperature.

21. The process of claim 8 in which said mixture

includes elemental sulfur and the temperature of said

brine leach is at the agglomeration temperature of sulfur.

22. The process of claim 19 in which lead is recovered

from the crystallized lead cWoride.

23. The process of claim 22 in which the lead is recovered

by hydrogen reduction of the crystallized lead

chloride.

24. The process of claim 23 in which the hydrogen

and the heat requirement for the endothermic reduction

of lead chloride are supplied by a partially combusted

hydrocarbon gas.

25. The process of claim 23 including supplying the

heat required for the endothermic reduction by heating

molten lead introduced into the reaction area.

26. A process for solubilizing a cWoride selected from

the group consisting of lead chloride and a combination

of silver and lead chlorides which comprises subjecting

the chloride to a brine leach at a temperature above the

normal boiling point under a pressure above atmospheric

sufficient to prevent boiling for a time not in

excess of 3 minutes sufficient to solubilize substantially

all of the chloride.

27. A process for solubilizing a cWoride selected from

the group consisting of lead cWoride and a combination

of silver and lead chlorides which comprise subjecting

the cWoride to a brine leach at a temperature above the

normal boiling point under a pressure above atmospheric

sufficient to prevent boiling for a sufficient time

to obtain a concentration of at least about 130 gplof

solubilized lead cWoride. • • • • •

0.022

0.015

0.020

0.D15

0.D18

0.017

0.023 .

0.018

0.012

0.013

0.010

0.112

0.026

0.019

0.013

0.012

0.010

0.009

In the third stage of the process of the present invention,

pulp from the pressure oxidation step is passed

through a surge tank to which a basic chemical may be

35 added for pH adjustment. It has been found that the

pulp from the pressure oxidation step should preferably

be cooled to a temperature below about 50· C. before it

is introduced to carbon-in-Ieach treatment. This cooling

may take place in a heat exchanger which recovers part

of the heat to be used upstream from the heat exchanger.

In addition to cooling the pulp, it may be

necessary to dilute the pulp before the carbon-in-Ieach

treatment wherein cyanide may be added to the first of

a plurality of mechanically agitated vessels in series in

which gold extraction from ore by cyanidation and

carbon absorption will proceed simultaneously. Pulp is

transferred continuously downstream through interstage

screens from a first vessel to the following vessel

in the series of vessels while activated granular charcoal

carbon is advanced from the last vessel toward the first

vessel. Fresh reactivated granular charcoal carbon is

added to the last stage and the loaded charcoal is withdrawn

from the first stage. Pulp leaving the last stage is

passed through an additional screen to scavenge some

attrited carbon before being discarded to a tailings pond

from which the tailings water may be recycled to the

slurry formation step.

When liquid from the tailings pond is recycled upstream

of the autoclaves, it has been found that free

cyanide in the recycled liquid unexpectedly decreases

the efficiency of the gold recovery of the present invention.

This is in contrast to conventional cyanidation

techniques in which the excess free cyanide would in

fact be utilized to help oxidize the slurry to be treated.

To overcome this unexpected problem, the free cyanide

must be removed by any suitable means.

The invention will be further illustrated in the following

example in which the pressure oxidation step is

20 Composite DC-I:

Composite DC-2:

Temp 0c.

180 50 0.028

200 ~ 0.020

215 ~ 0.017

225 ~ 0.016

249 ~ 0.021

250 ~ 0.016

150 100 0.070

180 ~ 0.031

180 ~ 0.027

215 ~ 0.021

225 ~ 0.011

Composite 2:

150 25 0.020

180 ~ 0.017

225 ~ 0.007

Composite DC-2:

150 0.060

180 0.023

200 0.014

215 0.014

225 0.010

All tests at 60 minutes and 40% solids.

In order to achieve adequate mixing without unnecessarily

diluting the slurry, it has been found preferable

to maintain the solids content of the slurry between

about 40% to about 50% and it is postulated that the

rate of reaction in the pressure oxidation stage is con- 40

trolled by the mass transport of oxygen to the solids'

surface. However, the selection of the proper type of

mixing equipment and the mixing speed utilized within

the autoclaves is deemed to be well within the scope of

one of ordinary skill in the art. 45

Another critical variable in the pressure oxidation

step is the residence time which the ore slurry spends in

the autoclaves. By way of example only, a summary of

tests evaluating the effect of autoclave time is given in

Table 7 and shown graphically in FIG. 2. The data 50

show a curve with minimum slope from approximately

30-90 minutes. Progressively higher tails assays result

from shorter times. Complete oxidation of the sulfides is

unnecessary; within an average 30-minute treatment

period found to be adequate, less than 50% of the sul- 55

fides may be oxidized. To maintain the oxygen pressure,

a bleed is required from the autoclaves as carbon dioxide

is evolved from the reaction of carbonates noted

above. Additionally, since the amount of oxygen consumed

will inevitably vary depending upon the refrac- 60

tory nature of the ore being treated, the oxygen partial

pressure must be monitored to assure it does not drop

below a minimum of about 10-25 psia. The partial pressure

of oxygen may be maintained by the introduction

of either pure oxygen, air or a mixture of both into the 65

autoclaves to ensure that an effective amount of the

substances which cause the ore to be refractory will be

oxidized in the pressure oxidation step.

4,552,589

described in greater detail. Thus, the operating conditions

in the autoclaves are as follows: a temperature of

about 2250 C., a retention time of about 30 minutes, an

oxygen overpressure· of about 25 psia, an agitation of

about 100 hp/5,OOO gallon unit, a pulp density of about 5

45% solids, a grind of about 80% passing through 325

mesh, a reagent addition of about 8 pounds NaOH/ton

ore, an oxygen supply (average) of about 45 lb/ton ore

and a steam supply (maximum) of about 250 1b/ton of

ore. At least six stages are utilized to avoid by-passing 10

inefficiency. This might be achieved through six vertical

units or one or two multistage horizontal autoclaves.

The gas bleed and oxygen input are controlled by manifolding

the bleed gas from each stage and exhausting to

control pressure, while injecting oxygen to maintain a 15

desired oxygen concentration in the atmosphere of each

vessel. The agitation is designed to reentrain gas at the

pulp surface. A flash heat exchange system is used to

recover heat, followed by steam injection to attain the

desired pulp temperature in the first stage. Bleed gas is 20

contacted with feed pulp to recover some heat as well

as to provide initial scrubbing. Although the pulp

would be alkaline, pitting corrosion could occur above

the pulp interface. Accordingly, a mild steel autoclave 25

would require at least a partial protective lining.

Having fully described the present invention, it will

be apparent from the above description and drawings

that various modifications in the process of the present

invention may be made within the scope of this inven- 30

tion. Therefore, this invention is not intended to be

limited except as may be required by the lawful scope of

the following claims.

What is claimed is:

1. A process for recovering gold from a refractory 35

ore slurry, said process comprising the steps of:

introducing the slurry to at least one agitated autoclave

maintained at an autoclave temperature in

excess of about 1500 C. and an autoclave oxygen

partial pressure in excess of about 10 psia; 40

agitating the slurry to allow oxygen to oxidize at least

part of the substances which cause the ore to be

refractory to form an oxidized slurry;

passing the oxidized slurry to a plurality of stages in

which the oxidized slurry is contacted with a cya- 45

nide and activated charcoal, the charcoal flowing

in countercurrent fashion with said oxidized slurry

whereby precious metals are transferred to the

activated charcoal; and

separating the activated charcoal from the oxidized 50

slurry.

2. The process as recited in claim 1 wherein the autoclave

temperature is maintained between about 1800 C.

to about 2250 C.

3. The process is recited in claims 1 or 2 wherein the 55

autoclave oxygen partial pressure is maintained at a

minimum between about 10 psia and about 25 psia.

4. A process as recited in claim 1 wherein a basic

compound, is added to the ore slurry to form an insoluble

species with free sulfate ions.

5. A process as recited in claim 4 wherein the basic

compound comprises sodium carbonate.

6. A process as recited in claim 1, comprising the

further step of:

cooling the oxidized slurry before the oxidized slurry 65

is contacted with the cyanide.

7. A process as recited in claim 1, comprising the

further step of:

raising the pH of the oxidized slurry before the oxidized

slurry is contacted with the cyanide.

8. A process as recited in claim 1, comprising the

further step of:

diluting the oxidized slurry before the oxidized slurry

is contacted with the cyanide.

9. A process as recited in claim 1 wherein the ore

slurry is maintained within the autoclave step from

approximately 30-90 minutes.

10. A process for recovering gold, said process comprising

the steps of:

forming an ore slurry from a refractory gold bearing

ore;

introducing the slurry to to at least one autoclave

which is maintained at an autoclave temperature in

excess of about 1500 C. and an autoclave oxygen

partial pressure maintained in excess of about 10

psia;

agitating the slurry sufficiently to oxidize at least part

of the substances which cause the ore to be refractory

to form an oxidized slurry;

passing the oxidized slurry to a plurality of stages in

which the oxidized slurry is contacted with a cyanide

and activated charcoal at a cyanidation temperature

less than about 500 C., said charcoal flowing

in countercurrent fashion with said oxidized

slurry whereby gold is transferred to the activated

charcoal; and

separating the activated charcoal from the oxidized

slurry.

11. The process as recited in claim 10 wherein the

autoclave temperature is maintained between about

1800 C. to about 2250 C.

12. The process as recited in claims 10 or 11 wherein

the autoclave oxygen partial pressure is maintained at

about 15 psia.

13. A process as recited in claim 10 wherein a basic

compound is added to the ore slurry to form an insoluble

species with free sulfate ions.

14. A process as recited in claim 13 wherein the basic

compound is selected from the group consisting of sodium

carbonate, sodium bicarbonate and sodium hydroxide.

15. A process as recited in claim 10 wherein the gold

in the ore exists in a finely disseminated form with an

average size of less than 10 microns.

16. A process as recited in claim 10 wherein the refractory

ore contains a sulfur bearing compound which

is at least partially oxidized in at least one autoclave.

17. A process as recited in claim 10 wherein the ore is

less than 60% amenable to standard cyanidation techniques.

18. A process as recited in claim 17 wherein more

than 85% of the gold existing in the ore slurry is transferred

to the activated charcoal.

19. A process as recited in claim 10 wherein the ore

slurry is maintained within at least autoclave for from

about ten minutes to about thirty minutes.

20. A process as recited in claim 10 wherein the activated

charcoal is separated from the oxidized slurry

from about 4 hours to about 8 hours after said activated

charcoal is contacted with the oxidized slurry.

21. A process for recovering gold, said process comprising

the steps of:

forming an ore slurry by grinding a refractory gold

bearing ore to form a slurry which is mixed with a

diluent stream;

4,552,589

introducing the slurry to at least one autoclave which

is maintained at an autoclave temperature in excess

of about 150· C. and an autoclave oxygen partial

pressure maintained in excess of about 10 psia;

agitating the slurry to oxidize substances which cause 5

the ore to be refractory to form an oxidized slurry;

cooling the oxidized slurry to a temperature less than

about 150· C.;

passing the oxidized slurry to a plurality of stages in

which the oxidized slurry is contacted with a cya- 10

nide and activated charcoal, the activated charcoal

flowing in countercurrent fashion with the oxidized

slurry whereby gold is transferred to the

activated charcoal; 15

separating the activated charcoal from the oxidized

slurry;

collecting a tailings stream from the oxidized slurry;

treating a water portion of the tailings stream to destroy

any free cyanide and form a treated reclaimed 20

water stream; and

recycling the treated reclaimed water stream to a step

ofthe process upstream from the step of passing the

oxidized slurry to a plurality of stages.

22. A process as recited in claim 21 wherein the

treated reclaimed water stream is combined with a feed

stream to form the diluent stream.

23. A process as recited in claim 21 wherein the oxidized

slurry is cooled by a heat exchanger which is

utilized to preheat the slurry.

24. A process as recited in claim 21, comprising the

further step of:

diluting the oxidized slurry to dethicken said oxidized

slurry before the oxidized slurry is contacted with

the cyanide.

25. A process as recited in claim 24 wherein the reclaimed

water stream is used to dilute the oxidized

slurry.

26. A process as recited in claim 21 wherein the ore

slurry is maintained within at least one autoclave from

about 10 minutes to less than about 60 minutes and the

carbon-in-leach residence time is from about 4 hours to

about 8 hours.

27. A process as recited in claim 26 wherein the ore

slurry is less than 60% amenable to standard cyanidation

techniques and more than 85% of the gold in said

ore slurry is transferred to the activated charcoal.

* * * * *

ginati} ��oe@s� ut-grid-align:none;text-autospace:none'>portion of said high pressure leachate of step (c) and at

least a portion of said second fraction of step (a), are

contacted in an atmospheric leach at atmospheric pressure

and at a temperature below about 100· C. to form 20

an atmospheric leach residue and an atmospheric leachate

prior to contact in the low pressure leach of step (d).

4. A method according to claim 3 further comprising

separating from said first fraction a coarser third frac- 25

tion, said third fraction containing more magnesium

than said first fraction and less than said second fraction,

and contacting at least a portion of said third fraction

and at least a portion of said atmospheric leachate in the

low pressure leach of step (d). 30

5. A method according to claim 1 further comprising

separating from said first fraction a coarser third fraction,

said third fraction containing more magnesium

than said first fraction and less than said second fraction,

and contacting at least a portion of said third fraction in 35

the low pressure leach of step (d).

6. A method according to claim 1 or claim 3, wherein

at least a portion ofsaid low pressure leachate is neutralized

by the addition of a neutralization agent selected

from the group consisting of alkali and alkaline earth 40

oxides and alkali and alkaline earth hydroxides to form

a neutralized low pressure leachate.

7. A method according to claim 1, or claim 2, or claim

3, or claim 4, wherein at least a portion of said low

pressure leach residue is recycled to said high pressure 45

leach.

8. A method according to claim 3 or claim 4, wherein

at least a portion of said low pressure leach residue is

recycled to said atmospheric leach.

9. A method according to claim 8 wherein a coarser 50

fraction of said atmospheric leach residue is separated

from said atmospheric leachate and the remainder of

said atmospheric leach residue, and further comprising

filtering said coarser fraction into a filtrate and a filter

cake from which chromite is recovered.

10. A method according to claim 9 wherein at least a

portion of said filtrate is recycled to said atmospheric

leach.