3,709,680 Process for removal of arsenic from sulfo-ore

Patent Number/Link:

United States Patent [19)

Holmes et ale

[ II]

[45)

3,709,680

Jan. 9, 1973

FOREIGN PATENTS OR APPLICATIONS

Primary Examiner-Allen B. Curtis

Attorney-March and Le Fever

1899 Great Britain 75/118

[54) PROCESS FOR REMOVAL OF

ARSENIC FROM SULFO·ORE

[75] Inventors: W. Church Holmes, Lake San Marcos,

Calif.; Enzo L. Coltrlnari, Arvada,

Colo.

[73] Assignee: Sunshine Mining Company, Kellog,

Idaho

[22] Filed: July 9,1971

[21] AppI. No.: 161,165

2,835,569

3,218,161

3,476,553

22,619

[57]

5/1958

11/1965

11/1969

Reynaud et al... 75/118

Kunda et ai. 75/118 X

Sebbaetai. 75/121 X

ABSTRACT

[52] U.S. CI 75/6, 75/101 R, 75/118,

75/121

[51 ] Int. CI. C21b 1/04, C22b 61/00

[58] Field ofSearch 75/101 R, 118, 121,6

[56] References Cited

UNITED STATES PATENTS

683,325 9/1901 PhiIlips , 75/118

726,294 4/1903 Hoyt... 75/118

A process and apparatus for the removal of arsenic

values from an ore concentrate includes the steps of

leaching the concentrate with a solution which dissolves

the arsenic, separating undissolved residue

therefrom, acidifying the arsenic pregnant solution to

precipitate arsenic and other mineral values as insoluble

salts and treating the arsenic barren solution to

regenerate sulfur and sodium values for recycle to the

process.

10 Claims, 2 Drawing Figures

/1"2 S SOLtlT/OH

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,#u TO FtI/i'THE/i' P/i'l:JCESS//Y4

HZs----+----

S02 TO RECYCLE

NA'... S Ttl R,tCYCLE

PATENTEOJAH' 91973 3.709.680

SHEET 1 OF 2

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PATENTEOJAH 9 1973

SHEET 2 OF 2

3.709,680

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INVENTORS

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3,709,680

The temperature of the concentrated sodium sulfide

solution used to dissolve the arsenic from the flotation

concentrate must be carefully controlled in order to

obtain the optimum results. The sodium sulfide solu-

5 tion normally contains from about 100 to about 600

grams per liter of sodium sulfide, sodium sulfide-sodium

hydroxide or the equivalent of the other mixtures.

The temperature of the leaching solution in order to

obtain the maximum results in an efficacious and rapid

10 manner should be within the range at sea level of from

about 75°C to the temperature just below boiling,

preferably from about 100° to about 205cC. It has been

found that optimum extraction of arsenic from the

flotation concentrate occurs if the leaching time is

15 maintained at one within a range of from about I to 20

h~urs, preferably from about 4 to 12 hours. Agitation is

preferably maintained to aid in insuring an adequate

contact between the leaching solution and the flotation

20 concentrate.

The major content of the concentrate remains in the

residue as the corresponding sulfides, and is separated

in that form for value recovery.

The hot leach mixture from the leaching operation is

25 diluted with preferably hot water prior to separation of

residues. The dilution prevents undesirable crystallization

of the salts with temperature drop and results in

more efficient and effective further processing.

The diluted mixture is preferably dewatered by a plu-

30 rality of thickeners, arranged for counter-current decantation,

the thickened slurry filtered and the filtrant

contains desirable mineral values. The filtrate is combined

with no decant from the thickening step and

passed to the next stage of the process.

As an example of the leaching step the following is

given.

A sample of a flotation concentrate containing 12.2

percent arsenic and weighing 100 grams was contacted

with a solution containing 390 grams of sodium sulfide

per liter of water. The solution was agitated for 8 hours

at a temperature of 100cC. After filtering and washing

with hot water, the residue assayed 0.5 percent arsenic.

2. Treatment of the Arsenic Pregnant Solution

The solution from the leaching step may contain, in

addition to the arsenic values, gold, antimony, mercury,

sodium and sulfur values which are recoverable and

which the process of this invention is designed to

recover. The arsenic removal and gold recovery is accomplished

in a series of steps which will now be

50 described.

A. Recovery of Gold Values

The ore concentrate described above contains approximately

1.55 ounces of gold concentrate. Although

leaching conditions may be controlled to regulate the

amount of gold that is dissolved, any gold present in the

arsenic pregnant solution may be removed by reducing

the pH of the solution to one within the range of from

about 9-11. This acidification step is preferably carried

out by contacting the solution with certain types of

acidifying agents as, for example, S02 gas. The contact

is carried out until the pH of the solution is at the

desired level, usually pH 10. At this point, any gold

present is precipitated, filtered, washed, dried for

recovery of the mineral values.

During the acidification step hydrogen sulfide is

evolved which is normally purified and recycled to the

process.

(approximate)

48.0%

19.0%

33.0%

PREPARATION OF CONCENTRATE

BRIEF DESCRIPTION OF THE INVENTION

PROCESS FOR REMOVAL OF ARSENIC FROM

SULFO·ORE

BRIEF STATEMENT OF THE INVENTION

The novel process of this invention may be logically

divided into a number of sequential steps as follows:

I. Leaching of the concentrate with a caustic solution;

2. Treatment of the arsenic pregnant solution; and

3. Treatment of the arsenic barren solution.

There are a number of processing steps involved in

each of the above-mentioned general classifications in

which conditions of the reaction are critical and must

be carefully controlled to obtain the optimum results.

These critical conditions will be particularly noted as

this description proceeds.

This invention relates to a process for the removal of

arsenic from materials in which it is contained. Particularly

the invention relates to a novel process for the

removal of arsenic from mixed ore mineral concentrates

containing a minor proportion thereof.

More particularly, the invention relates to an improved

process for the removal and/or recovery of arsenic

from ore concentrates which may contain mixed

values of copper-gold-sulfur-arsenic-antimony and

which involves the steps of leaching the arsenic values

from the ore concentrate, causing arsenic compounds

to precipitate from the solution and recovering values

such as gold, antimony, etc., from the leach solution.

Arsenic represents an undesirable contaminant in

smelting operations particularly in a number of coppercontaining

minerals. Exemplary of such minerals are

the sulfo-arsenic salts of copper, sulfo-antimony salts of

copper and mixtures such as tennantite, luzonite, 40

enargite, tetrahedrite, famantinite and the like. One

particularly valuable ore mineral containing components

in which the arsenic content of approximately

19 percent is especially undesirable is enargite having 45

approximately the following composition:

Copper

Arsenic

Sulfur

This particular mineral is often found to include

minor amounts of gold, silver, antimony, mercury,

bismuth. tin, zinc, lead and other metals.

Generally speaking, the concentrate is prepared by a 55

selective flotation operation, as is well known in the art.

I. Leaching of the Concentrate with a Caustic Solution

In accordance with the inventive process the concentrate

is leached with a hot concentrated caustic solution,

preferably of sodium sulfide, although it is possi- 60

ble to use sodium hydroxide or mixtures of sodium sulfide

and sodium hydroxide, or sodium hydroxide and

sulfur. Since the undissolved residue from the leaching

step contains valuable mineral valves, the undissolved 65

solids from the leaching step are settled, filtered,

washed and dried. They are then removed from the

process for metal recovery by conventional techniques.

3,709,680

It is to be understood that the apparatus which is

hereinafter described is only illustrative of one type of

apparatus which may be used to obtain the desired

results and the specification thereof is simply to

facilitate an understanding of the various steps involved.

This exemplary iIlustration is not to be considered

as a limitation of any kind on the apparatus as

defined in the appended claims.

Turning now to the drawings, and with particular

reference to FIG. 2, there is shown in the block diagram

form an arrangement of apparatus wherein the

feed to the process, an ore concentrate, is fed to

leaching vessel 10, which may be equipped with an

agitator driven by a motor through line 16. It will be appreciated,

of course, that for handling solid materials,

line 16 may be a screw or other type of conveyor means

and the latching vessel wiII be equipped with a hopper

as a feed inlet. If desired the leaching zone may be a

plurality of leaching vessels and may be provided with

means for temperature control as known to the art.

Hot sodium sulfide solution is added to leaching vessel

10 through line 18. After the desired leaching time

the slurry is passed to a series of thickening vessels arranged

for countercurrent washing, designated here by

numeral 20. Each thickener may be provided with a

scraper or rake to move the thickened slutty or residue

toward the center of the vessel bottom.

The leach solution is introduced to the thickening

zone by means of line 38. Water at the desired temperature

is introduced to the thickener 20, such that

the slurry of leaching solution and solids passes

counter-currently thereto.

Thickened solids are withdrawn from settler 20

35 through line 48 and passed to filter 50. The filtrate is

recycled to the process through line 52. Separated

solids, which represent the mineral values in the feed

concentrate, are removed from the filter through line

54 and set for further processing.

The counter-current wash or decant from the settlers

or thickening zone is withdrawn through line 56 and introduced

into first acidification vessel 62. This vessel,

which may be any of the known liquid-gas contacting

vessels known to the art, is equipped with inlet line 62

and a gas distribution means for bubbling an acidifying

gas, such as S02, through a column of the arsenic

pregnant solution in the vessel.

Acidification is continued in this vessel until a pH of

within a range of from about 9-11, preferably 10, is

reached. The solution with precipitated gold values,

normal1y in the form of its sulfide, is then withdrawn

through line 66, for example, and introduced to

thickener 68 which may be equipped with a rake driven

by a motor.

In this thickener the solids are allowed to settle,

gathered by the rake and withdrawn through line 74 to

filter 76. Separated solids which contain the

precipitated gold sulfides, are removed from the filter

through line 77 for processing to recover the gold con-

60 tent thereof.

The arsenic pregnant solution from thickener 68,

with gold content removed, and the filtrate from filter

76, through lines 78 and 80, is passed into a second

65 acidification vessel 82. In this liquid-gas contacting vessel,

as in vessel 62, an acidifying gas, such as S02is introduced

at the bottom, through line 86 and a sparger,

so that intimate liquid gas contact is obtained and

DETAILED DESCRIPTION OF THE APPARATUS

OF THE INVENTION

The invention is more clearly explained by reference

to the following drawings in which:

FIG. 1 represents a block diagram of the flow sheet

of one embodiment of the invention; and

FIG. 2 is a process block diagram of one form of apparatus

useful for carrying out the process.

B. Removal of Arsenic

The arsenic pregnant solution which has been

treated to recover dissolved gold is now subjected to a

second acidification step. Certain types of acidifying

agents are again used, such as S02 gas, and the solution 5

is contacted until the pH is reduced to one within a

range of about 2 to 6, preferably about 3 to 4. The H2S

generated by this acidification step is cycled for use at a

subsequent step in the process.

At a pH of 3 and at a temperature of 25° to 50°C, 10

preferably below about 50°C, the arsenic values in the

arsenic pregnant solution precipitate in the form of sulfides,

and the solid pulp is filtered and washed. The

solids from the filtration step comprise essentially the

arsenic sulfides and are normally discarded. These are 15

normally the tri- or pentasulfides.

3. Treatment of the Arsenic Barren Solution

The arsenic barren solution contains sodium and sulfur

values which are recovered and recycled for use.

In summary, this recovery involves evaporation of 20

the solution to crystalIize these values in the form of

salts, roasting the salts to recover sulfur in the form of

S02 gas and thereby oxidizing the sodium salts to

sulfates and thereafter reducing the calcine with a

source of carbon to convert the sodium values to sodi- 25

urn sulfide which is dissolved and recycled to the

leaching step.

A. Evaporation Step

Evaporation of the arsenic barren solution is carried

out in evaporation equipment known to the art and 30

yields a variety of sodium-sulfur compounds.

If desired a portion of the sulfur obtained during the

evaporation step may be removed prior to passing the

material to the oxidation roast.

B. The Oxidation Roast

The solids from the evaporation step containing the

sulfur and sodium values are conveyed to a roasting

zone where they are subjected to temperatures in the

order of 150° to 600°C, preferably 400° to 500°C. The

roasting is conducted in an oxidizing atmosphere nor- 40

mally in the form of air. H2S obtained from the previous

step may also be introduced. The effluent gases from

the roasting step contain approximately 13% S02 by

volume, which are recycled to the acidification step.

The calcine obtained from this roast is passed to the 45

next step in the process.

C. The Reducing Roast

The calcine from the first roasting step is preferably

admixed with a source of carbon, normally in the form

of a pulverized coal. After subjecting the calcine to an 50

adequate temperature for the desired reaction time, the

sodium sulfate contained in the calcine is reduced to

sodium sulfide.

Product from this second roasting or furnacing step

contains about 75% Na2S, the remainder being ash, un- 55

burned coal, and impurities. The product is then mixed

with water and the dissolved sodium sulfide is recycled

to the process.

3,709,680

7 8

furnacing said solid residue at a temperature of from

about 850° to I,200°C in the presence of carbon to

recover solid sodium sulfide therefrom;

dissolving solid sodium sulfide in water to form a

solution thereof; and 5

recycling sodium sulfide solution to the process.

* * * * *

-size:7f�G;o��0�:"Times New Roman","serif";mso-fareast-font-family: HiddenHorzOCR'>Molybdenum _

Potassium_ •• _

Silicon_ • _

Sodium • •

Tin • _

Others_ • _

I Not dried.

Tables 1 and 2 show that up to 98% of the molybdenum

contained in the original feed samples was recovered by

the process. Tables 1, 2, 5 and 6 show that up to 99%

of rhenium contained in the original samples was recovered

by the process. The final products obtained were

substantially free of impurities derived from the feed

solution.

The invention described provides an effective and

economical method for almost 1<00% recovery of molybdenum

and rhenium from solutions in which they are

present together. Although the solutions from which

the recoveries are made are ordinarily scrubber solutions

resulting from roasting of molybdenite concentrates, the

invention is not restricted to recovery of the metals from

this type solution.

What is claimed is:

1. A process for recovering molybdenum and rhenium

values from solutions in which they are present together

with other metal ion impurities which comprises:

(a) extracting the rhenium and molybdenum values

from the solution with a liquid water insoluble amine

ion exchange agent;

(b) stripping the rhenium and molybdenum values from

the loaded agent of (a) with a basic solution of an

ammonium compound;

from the ship solution of (b) by adjusting

the pH of the strip solution to about 2.0-3.5;

157 15

5.9

5.4

3.6o

1.1

0.6

0.5o

TABLE 5

Data for Extraction of Rhenium and Rejection of Molybdenum by Pyridine from 6 M Sodium Hydroxide

Solution

Molybdenum Rhenium

Grams per liter Grams per liter

Loaded Percent Loaded Percent

Test No. Feed solvent Raffinate extracted E, alo Feed solvent Raffinate extracted E,a!o

L __ e __•• __ 14.5 0.35 11.4 5 0.03 3.05 1.7 0.010 99.6 170

32_____________.:_c__. 17.1 0 18.4 0 0 1.42 1.2 0.008 99.5 155

14.5 0.09 9.9 5 0.009 3.05 0.40 0.003 99.8 133

4______ •• _. 14.5 0.94 10.8 20 0.09 3.05 3.3 0.03 99.0 110

5___ •• ____ • 14.5 0.07 10.5 3 0.007 3.05 0.84 0.01 99.5 84 6____ • ___ •• 18.4 ___ • __ • ___ • ___• _______ ••••c_._.__________ • 1.5 1.48 0.018 99.1 82 7_.________

40.9 0 41.1 0 0 16.8 8.7 0.107 99.3 81

8__ • _______ 18.4 • ____________• __ • _________________________

9____ • _____ 1.5 1.48 0.025 98.7 59

40.9 0 45.0 0 0 16.8 17.6 0.326 98.0 54

10_________ 40.9 0 46.1 0 0 16.8 44.7 5.26 69 8.5

Portions of rhenium have been recovered from the extract

by both of these procedures. In the first, the rhenium 70

was precipitated from the distillation bottoms as the

rhenium heptasulfide. In the second, the rhenium was reduced

to metal with hydrogen. The potassium and sodium

salts were removed by leaching with water and dilute hydrochloric

acid. Table 6 shows the quality of the rhenium 75

(d) recovering the crystallized ammonium tetramolybdate

of (c) followed by recovery of molybdenum

values therefrom;

(e) extracting rhenium 'Values from the mother liquor

of (d) with a liquid water insoluble amine ion exchange

agent;

3,681,016

25 HERBERT T. CARl1ER, Primary Examiner

u.s. Cl. X.R.

ing the pH of rthe solution' to a pH between about

2.5-3.5;

(d) separating the crystallized ammonium tetramolybdate

from the mother liquor of (c);

(e) extracting rhenium values from the mother liquor

of (d) with a liquid water insoluble amine ion exchange

agent;

(f) stripping rhenium values from the loaded agent of

(e) with sodium hydroxide;

(g) recovering rhenium values from the strip solution

of (f) by extracting with pyridine; and

(h) recovering rhenium from the pyridine extractant

by distilling off the pyridine.

References Cited

UNITED STATES PATENTS

7/10969 Litz , 23-15 W

7/1969 !Platzke et al. 23-15W

3/1959 Zimmerley et al. 23-18 X

7/1960 Zimmerley et al. 23-24

4/1966 Churchward , 23~15 W

211970 Ziegenbaly et al. 23-23 X

1/1971 Proter et al. 23-22

3,455,'677

3,45:8,277

2,876,065

2,945,743

3,244,475

3,495,934

3,558,268

23-23, 24 R, 51 R

(f) stripping the loaded agent of (e) with an alkali

metal hydroxide;

(g) extracting rhenium values from the strip solution

of (f) with pyridine or pyridine derivative; and

(h) recovering rhenium from the pyridine extractant 5

by distilling off the pyridine.

2. The process of claim 1 in which metal ion impurities

are removed from the strip solution of .(b) before

crystallizing ammonium tetramolybdate in (c).

3. The process of claim 1 in which the anion exchange 10

agent in (a) is a tertiary amine ion exchange resin and the

stripping solution of (b) is ammonium hydroxide.

4. A process for recovering molybdenum and rhenium

values from pregnant acid leach solutions containing these

values together with other metal impurities and derived 15

from dusts and flue gases resulting from roasting relatively

impure molybdenite concentrate, said process comprising:

(a) extracting molybdenum and rhenium values from

the pregnant acid solution with a liquid water in- 20

soluble amine ion exchange agent;

(b) stripping the molybdenum and rhenium values

from the exchange resin with ammonium hydroxide

solution to form a strip solution containing the molybdenum

as ammonium molybdate and the rhenium

as ammonium perrhenate;

in (b) as ammonium tetramolybdate by adjust