4,879,022 Ore flotation process and use of mixed hydrocarbyl dithiophosphoric acids and salts thereof

Patent Number/Link:

United States Patent [19]

Clark et at

[11] Patent Number:

[45] Date of Patent:

4,879,022

Nov. 7, 1989

[54] ORE FLOTATION PROCESS AND USE OF

MIXED HYDROCARBYL

DITHIOPHOSPHORIC ACIDS AND SALTS

THEREOF

OTHER PUBLICAnONS

Froth Flotation-50th Anniversary vol. Fuerstenaw

(Editor) Published by AIMM & PE Inc. 1962, pp.

420-422.

[75] Inventors: Alan C. Clark, Mentor; Edward P.

Richards, Chagrin Falls, both of

Ohio; Douglas R. Shaw, Lakewood,

Colo.

[73] Assignee: The Lubrizo1 Corporation, Wickliffe,

Ohio

(List continued on next page.)

Primary Examiner-Kenneth M. Schor

Assistant Examiner-Thomas M. Lithgow

Attorney, Agent, or Firm-Robert A. Franks; Frederick

D. Hunter; Forrest L. Collins

[21] Appl. No.: 300,694

Related U.S. Application Data

[63] Continuation of Ser. No. 72,809, Jul. 14, 1987, abandoned.

[51] Int. C1.4 B03D 1/02

[52] U.S. Cl. 209/166; 252/61;

241/24; 423/26; 209/167

[58] Field of Search 209/166, 167; 252/61;

75/2; 423/26; 241/24

The present invention relates to an improved process

for beneficiating an ore containing sulfide materials

with selective rejection of pyrite, pyrrhotite and other

metals and gangue. In particular, the process is useful

for beneficiating ores and recovering copper from said

ores. In one embodiment the process comprises the

steps of

(A) grinding the ore to an appropriate size range;

(B) preparing a slurry comprising

(B-1) said ground ore;

(B-2) at least one collector which is a water-dispersible

or soluble dihydrocarbyldithiodiphosphoric

acid or salt having the formula

[57] ABSTRACT

[22] Filed: Jan. 19, 1989

References Cited

U.S. PATENT DOCUMENTS

1,377,189 5/1921 Dosenbach 209/167

1,486,297 3/1924 Pallanch 209/167

1,593,232 7/1926 Whitworth 209/166

1,678,259 7/1928 Martin 209/167

1,873,115 8/1932 Derby 209/166

2,012,830 8/1935 Ralston 209/167

2,038,400 4/1936 Whitworth 209/166

2,063,629 12/1936 Su1zberg 209/166

2,206,284 7/1940 Jayne, Jr 252/9

2,849,475 8/1958 Capenter 252/61

2,919,025 12/1959 Booth et al. 209/166

3,086,653 4/1963 Booth 209/166

(List continued on next page.)

(I)

ECTOR

wherein RI and R2 are different hydrocarbyl

groups containing up to about 12 carbon atoms, n is

an integer equal to the valence of X and Xn+ is a

dissociating cation; and

(B-3) water; .

pH of about 5.5 to about 7.5;

(D) subjecting the conditioned slurry to froth flotation

to produce a froth containing a metal rougher concentrate;

(E) collecting said froth; and

(F) recovering the metal rougher concentrate containing

the desired metal values.

34 Claims, 1 Drawing Sheet

COPPER CLEANER

TAILINGS

ORE

FOREIGN PATENT DOCUMENTS

499430 4/1979 Australia.

[56]

COPPER

CLEANER

CONCENTRATE

4,879,022

Page 2

U.S. PATENT DOCUMENTS

3,220,551 1111965 Moyer 209/167

3,570,772 3/1971 Booth et al. 241124

4,040,950 8/1977 Zipperian 209/166

4,283,017 8/1981 Coale et al. 241124

4,460,459 7/1984 Shaw et al. 209/9

4,530,758 7/1985 Tibbals 209/166

4,684,459 8/1987 Klimpel 209/166

4,699,712 10/1987 Unger 209/166

OTHER PUBLICAnONS

"Organic Chemistry"by Morrison and Boyd, 3rd Edition

@1973 Allyn and Bacon Inc. p. 338.

The Separation by Flotation of Copper-Lead-Zinc

Sulphides by B. A. Willis. Minning Magazine 150, Jan.

1984.

Ullmanns Encyklopadie Der Technischen Chemie,

Weinheim 1979.

The Broken Hill Concentrator of Black Mountain Mineral

Development Co. (Pty.) Ltd., South Africa, Paper

Presented at the Complex Sufide Ores Conference, The

Institute of Mining and Metallurgy, Rome, Italy, Oct.

5-8, 1980.

Metallurgical Development at Woodlawn Mines, Australia,

Paper Presented at the Complex Sulfide Ores

Conference, The Institute of Mining and Metallurgy,

Rome, Italy, Oct. 5-8, 1980.

u.s. Patent Nov. 7, 1989 4,879,022

ORE

PRIMARY

GRIND

SLURRY H2O

PREPARATION COLLECTOR

CONDITIONER S02-

COPPER ROUGHER

ROUGHER TAILINGS

ROUGHER CONCENTRATE

REGRIND

COPPER COPPER CLEANER

CLEANER TAILINGS

COPPER

CLEANER

CONCENTRATE

FIG. I

4,879,022

(I)

SUMMARY OF THE INVENTION

wherein Rl and R2 are different hydrocarbyl

groups containing up to about 18 carbon atoms, n is

an integer equal to the valence of X and xn+ is a

dissociating cation;

(B-3) water; and optionally

(B-4) a water-soluble inorganic base;

pH of about 5.5 to about 7.5;

The present invention relates to an improved process

for beneficiating an ore containing sulfide materials

with selective rejection of pyrite, pyrrhotite and other

minerals and gangue. In particular, the process is useful

for beneficiating ores and recovering metals such as

copper, lead, zinc, etc., from said ores. In one embodiment

the process comprises the steps of

(A) grinding the ore to an appropriate size range;

(B) preparing a slurry comprising

(B-1) said ground ore;

(B-2) at least one collector which is a water-dispersible

or soluble dihydrocarbyldithiodiphosphoric

acid or salt having the formula

with phosphorus and sulfur generally as P2SS. The acid

obtained in this manner can then be neutralized to form

a salt which is stable yet soluble in water.

U.S. Pat. No. 3,086,653 describes aqueous solutions of

5 alkali and alkaline earth metal salts of phosphoorganic

compounds useful as promoters or collectors in froth

flotation of sulfide ores. The phosphoorganic compounds

are neutralized P2SS alkanol reaction products.

Although single alcohols are normally used in the reac-

10 tion, the patentees disclose that mixtures of isomers of

the same alcohol, and mixtures of different alcohols

may be utilized as starting materials in the preparation

of the phosphorus compound, and the resulting acidic

products can be readily neutralized to form stable solu-

15 tions which are useful as flotation agents.

U.S. Pat. No. 3,570,772 describes the use of di(4,5carbon

branched primary alkyl) dithiophosphate promoters

for the flotation of copper middlings. The 4 and

5 carbon alcohols used as starting materials may be

either single alcohols or mixtures of alcohols.

Procedures for the selective flotation of copper from

copper sulfide ores wherein a slurry of ore and water is

prepared and sulfurous acid is added to the slurry to

condition the slurry prior to the froth flotation step

have been discussed in, for example, U.S. Pat. Nos.

4,283,017 and 4,460,459. Generally, the pulp is conditioned

with sulfur dioxide as sulfurous acid under intense

aeration. This conditioning of the slurry enhances

the promotion and flotation rate of copper in the subsequent

flotation step. Generally, the amount of sulfur

dioxide added ranges from about 1 to 5 pounds of sulfur

dioxide per ton of ore. In U.S. Pat. No. 4,283,017, the

desirable pH of the conditioned slurry is reported to be

between about 5 and about 6.5, and preferably between

5.5 and 6.0. In U.S. Pat. No. 4,460,459, the pH of the

conditioned slurry is reported as being from about 6.5 to

6.8.

BACKGROUND OF THE INVENTION

ORE FLOTATION PROCESS AND USE OF MIXED

HYDROCARBYL DITHIOPHOSPHORIC ACIDS

AND SALTS THEREOF

This is a continuation of co-pending application Ser.

No. 072,809, filed on July 14, 1987, now abandoned.

TECHNICAL FIELD OF THE INVENTION

This invention relates to froth flotation processes for

the recovery of metal values from metal sulfide ores.

More particularly, it relates to the use of improved

collectors for beneficiating mineral values comprising

hydrocarbyl dithiophosphoric acids or salts derived

from a mixture of alcohols.

Froth flotation is one of the most widely used process

for beneficiating ores containing valuable minerals. It is

especially useful for separating fmely ground valuable 20

minerals from their associated gangue or for separating

valuable minerals from one another. The process is

based on the affinity of suitably prepared mineral surfaces

for air bubbles. In froth flotation, a froth or a foam

is formed by introducing air into an agitated pulp of the 25

finely ground ore in water containing a frothing or

foaming agent. A main advantage of separation by froth

flotation is that it is a relatively efficient operation at a

substantially lower cost than many other processes.

It is common practice to include in the flotation pro- 30

cess, one or more reagents called collectors or promoters

that impart selective hydrophobicity to the valuable

mineral that is to be separated from the other minerals.

It has been suggested that the flotation separation of one

mineral species from another depends upon the relative 35

wettability of mineral surfaces by water. Many types of

compounds have been suggested and used as collectors

in froth flotation processes for the recovery of metal

values. Examples of such types of collectors include the

xanthates, xanthate esters, dithiophosphates, dithiocar- 40

bamates, trithiocarbonates, mercaptans and thionocarbonates.

Xanthates and dithiophosphates have been

employed extensively as sulfide collectors in froth flotation

of base metal sulfide ores. One of the problems

associated with these conventional sulfide collectors is 45

that at pH's below 11.0, poor rejection of pyrite or

pyrrhotite is obtained. In addition, as the pH decreases,

the collecting power of the sulfide collectors also decreases

rendering them unsuitable for flotation in mildly

alkaline, neutral or acid environments. Suggestions 50

have been made in the art for modifications of the xanthates

and dithiocarbamates for improving their utility

as sulfide collectors in a variety of froth flotation processes.

Dialkyldithiophosphoric acids and salts thereof such 55

as the sodium, potassium, calcium or ammonium salts

have been utilized as promoters or collectors in the

beneficiation of mineral-bearing ores by flotation for

many years. Early references to these compounds and

their use as flotation promoters may be found in, for 60

example, U.S. Pat. Nos. 1,593,232 and 2,038,400. Ammonium

salt solutions of the dithiophosphoric acids are

disclosed as useful in U.S. Pat. No. 2,206,284, and hydrollyzed

compounds are disclosed as useful in U.S.

Pat. No. 2,919,025. 65

The diesters of dithiophosphoric acids utilized as

flotation promoters and collectors for sulfide and precious

metal ores are obtained by reacting an alcohol

4,879,022

(D) subjecting the conditioned slurry to froth flotation

to produce a froth containing a metal rougher concentrate;

(E) collecting said froth; and

(F) recovering the metal rougher concentrate contain- 5

ing the desired metal values.

BRIEF DESCRIPTION OF THE DRAWING

The sole FIGURE is a block-diagram representation

of a flow sheet of the flotation process according to the 10

present invention.

DETAILED DESCRIPTION OF THE

INVENTION

The froth flotation process of the present invention is 15

useful to beneficiate sulfide mineral and metal values

from sulfide ores including, for example, copper, lead,

zinc, nickel, and cobalt. Lead can be beneficiated from

minerals such as galena (PbS) and zinc can be beneficiated

from minerals such as sphalerite (ZnS), both of 20

which can be found in Central Missouri deposits. Cobalt-

nickel sulfide ores such as siegenite or linnalite also

available from Mississippi Valley deposits can be beneficiated

in accordance with this invention. The copper

sulfide minerals which can be beneficiated in accor- 25

dance with this invention are primarily chalcopyrites

(CuFeS2). The invention is useful particularly in beneficiating

the complex copper sulfide minerals such as

obtained from the Cerro, Colorado mines, central and

eastern Canada (Kidd Creek mine, New Brunswick 30

mines, etc.) Australia, Spain and South Africa. The

complex sulfide ores contain large amounts of pyrite,

(and other iron sulfides generally are relatively to beneficiate.

In the following description of the invention, how- 35

ever, comments primarily will be directed toward the

beneficiation and recovery of copper, and it is intended

that such discussion shall also apply to the other aboveidentified

minerals. The process of the present invention

has been found to be particularly useful in beneficiating 40

complex copper sulfide ores such as the Rio Tinto Minera

Cerro Colorado copper-pyrite ores.

The ores which are treated in accordance with the

process of the present invention must be reduced in size

to provide ore particles of flotation size. As is apparent 45

to those skilled in the art, the particle size to which an

ore must be reduced in order to liberate mineral values

from associated gangue and non-value metals will vary

from ore to ore and depends upon several factors, such

as, for example, the geometry of the mineral deposits 50

within the ore, e.g., striations, agglomerations, etc. Generally,

suitable particle sizes are minus 10 mesh (Tyler)

with 50% or more passing 200 mesh. The size reduction

of the ores may be performed in accordance with any

method known to those skilled in the art. For example, 55

the ore can be crushed to about minus 10 mesh size

followed by wet grinding in a steel ball mill to specified

mesh size ranges. Alternatively, pebble milling may be

used. The procedure used in reducing the particle size

of the ore is not critical to the method of this invention 60

so long as particles of effective flotation size are provided.

Water is be added to the grinding mill to facilitate the

size reduction and to provide an aqueous pulp or slurry.

The amount of water contained in the grinding mill be 65

varied depending on the desired solid content of the

pulp or slurry obtained from the grinding mill. Conditioning

agents as known in the art may be added to the

grinding mill prior to or during the grinding of crude

ore. Optionally, water-soluble inorganic bases and/or

collectors also can be included in the grinding mill.

In accordance with the process of the present invention,

an aqueous slurry is prepared containing the

ground ore and (B-2) at least one collector which is a

water-dispersible or soluble dihydrocarbyl dithiophosphoric

acid or salt having the formula

[

Ria ] ~P(S)S- xn+

R20

wherein Rl and R2 are different hydrocarbyl groups

containing up to about 18 carbon atoms, n is an integer

equal to the valence of X, and xn+ is a dissociating

cation. The amount of collector (B-2) included in the

slurry will depend upon a number of factors including

the nature of the ore, the size of the ore, etc. In general,

amounts of from about 0.01 to about 0.2 pound of collector

(B-2) may be used in the process of this invention

per ton of ore.

As used in this specification and the appended claims,

the term "ton" refers to a short ton, e.g., 2000 pounds.

The terms "hydrocarbyl" or "hydrocarbon-based" denote

a group having a carbon atom directly attached to

the remainder of the molecule and having predominantly

hydrocarbon character within the context of this

invention. Such groups include the following:.

(1) Hydrocarbon groups; that is, aliphatic, (e.g., alkyl

or alkenyl), alicyclic (e.g., cycloalkyl or cycloalkenyl),

aromatic, aliphatic- and alicyclic-substituted aromatic,

aromatic-substituted aliphatic and alicyclic groups, and

the like, as well as cyclic groups wherein the ring is

completed through another portion of the molecule

(that is, any two indicated substituents may together

form an alicyclic group). Such groups are known to

those skilled in the art. Examples include methyl, ethyl,

octyl, decyl, octadecyl, cyclohexyl, phenyl, etc.

(2) Substituted hydrocarbon groups; that is, groups

containing non-hydrocarbon substituents which, in the

context of this invention, do not alter the predominantly

hydrocarbon character of the group. Those skilled in

the art will be aware of suitable substituents. Examples

include halo, hydroxy, nitro, cyano, alkoxy, acyl, etc.

(3) Hetero groups; that is, groups which, while predominantly

hydrocarbon in character within the context

of this invention, contain atoms other than carbon

in a chain or ring otherwise composed of carbon atoms.

Suitable hetero atoms will be apparent to those skilled

in the art and include, for example, nitrogen, oxygen

and sulfur.

In general, no more than about three substituents or

hetero atoms, and preferably no more than one, will be

present for each 10 carbon atoms in the hydrocarbyl

group.

The hydrocarbyl groups Rl and R2 may be different

aliphatic, different aromatic, and/or mixtures of aliphatic

and aromatic groups containing up to about 18

carbon atoms. More generally, the alkyl groups will

contain from about 2 to about 12 carbon atoms, and the

aryl groups will contain from about 6 to about 18 carbon

atoms. Thus, in one embodiment, R1 and R2 are

different aliphatic groups; in a second embodiment, R1

and R2 are different aromatic groups, and in a third

4,879,022

EXAMPLE 4

A mixture of 2945 parts (24 equivalents) of Cresylic

Acid 57 (Merichem) and 1152 parts (6 equivalents) of

heptylphenol is heated to 105° C. under a nitrogen atmosphere

whereupon 1665 parts (15 equivalents) of

phosphorus pentasulfide are added in portions over a

period of 3 hours while maintaining the temperature of

the mixture between about 115°_120° C. The mixture is

maintained at this temperature for an additional 1.5

hours upon completion of addition of the phosphorus

pentasulfide and then cooled to room temperature. The

EXAMPLE 1

To 804 parts of a mixture of 6.5 moles of isobutyl

alcohol and 3.5 moles of mixed primary amyl alcohols

(65%w n-amyl and 35%w 2-methyl-l-butanol) is prepared,

and there are added 555 parts (2.5 moles) of

phosphorus pentasulfide while maintaining the reaction

temperature between about 104°-107° C. After all of the

phosphorus pentasulfide is added, the mixture is heated

for an additional period to insure completion of the

reaction and filtered. The filtrate is the desired phosphorodithioic

acid which contains about 11.2% phosphorus

and 22.0% sulfur.

EXAMPLE 2

The general procedure of Example 1 is repeated except

that the alcohol mixture reacted with phosphorus

pentasulfide comprises 40 mole percent of isopropyl

alcohol and 60 mole percent of 4-methyl-s-amyl alcohol.

The phosphorodithioic acid prepared in this manner

contains about 10.6% of phosphorus.

EXAMPLE 3

phenols. The aliphatic alcohols containing from about 4

to 6 carbon atoms are particularly useful in preparing

the dithiophosphoric acids and salts, etc.

Typical mixtures of alcohols and phenols which can

be used in the preparation of dithiophosphoric acids and

salts of Formula I include: isobutyl and n-amyl alcohols;

sec-butyl and n-amyl alcohols; propyl and n-hexyl alcohols;

isobutyl alcohol, n-amyl alcohol and 2-methyl-lbutanol;

phenol and n-amyl alcohol; phenol and cresol,

etc.

The phosphorodithioic acids and salts useful as collectors

in the process of the present invention are exemplified

by the acids and salts prepared in the following

examples. Unless otherwise indicated in the following

examples or elsewhere in the specification and claims,

all parts and percentages are by weight and all temperatures

are in degrees centigrade.

A mixture of 246 parts (2 equivalents) of Cresylic

Acid 33 (a mixture of mono-, di- and tri-substituted

alkyl phenols containing from 1to 3 carbon atoms in the

alkyl group commercially available from Merichem

Company of Houston, Tex.), 260 parts (2 equivalents) of

isooctyl alcohol and 14 parts of caprolactam is heated to

55° C. under a nitrogen atmosphere. Phosphorus pentasulfide

(222 parts, 2 equivalents) is added in portions

50 over a period of one hour while maintaining the temperature

at about 78° C. The mixture is maintained at this

temperature for an additional hour until completion of

the phosphorus pentasulfide addition and then cooled to

room temperature. The reaction mixture is filtered

through a filter aid, and the filtrate is the desired phosphorodithioic

acid.

embodiment, RI may be an aliphatic group and R2 an

aromatic group.

As noted, xn+ may be any dissociating cation, and in

one embodiment X is hydrogen, an ammonium group,

an alkali metal or an alkaline earth metal. Water-soluble 5

collectors generally are preferred, and thus, X normally

is an ammonium group, an alkali metal or certain Group

II metals. The alkali metals, sodium and potassium are

particularly preferred.

The dihydrocarbyldithiophosphoric acids and salts 10

represented by Formula I are known compounds and

may be prepared by the reaction of a mixture of hydroxy-

containing organic compounds such as alcohols

and phenols with a phosphorus sulfide such as P2SS.

The dithiophosphoric acids generally are prepared by 15

reacting from about 3 to 5 moles, more generally 4

moles of the hydroxy-containing organic compound

(alcohol or phenol) with one mole of phosphorus pentasulfide

in an inert atmosphere at temperatures from

about 50° C. to about 200° C. with the evolution of 20

hydrogen sulfide. The reaction normally is completed

in about 1to 3 hours. The salts of the phosphorodithioic

acids can be prepared also by techniques well known to

those in the art including the reaction of the dithiophosphoric

acid with ammonia, and various derivatives of 25

alkali and Group II metals such as the oxides, hydroxides,

etc. The formation of the salt typically is carried

out in the presence of a diluent (e.g., alcohol, water, or

diluent oil).

The composition of the phosphorodithioic acid ob- 30

tained by the reaction of a mixture of hydroxy-containing

organic compounds with phosphorus pentasulfide is

actually a statistical mixture of phosphorodithioic acids

wherein, with reference to Formula I derived from a

mixture of two hydroxy compounds, R10H and R20H, 35

RI and R2in one of the acids are different hydrocarbyl

groups derived from the different alcohols, RI and R2in

a second phosphorodithioic acid are identical and derived

from one of the alcohols, and RI and R2in a third

phosphorodithioic acid are identical but derived from 40

the second alcohol of the alcohol mixture. In the present

invention it is preferred to select the amount of the

two or more hydroxy compounds reacted with P2PS to

result in a mixture in which the predominating dithiophosphoric

acid is the acid (or acids) containing two 45

different hydrocarbyl groups. In the following Examples

1-4, the product is a statistical mixture of at least

three phosphorodithioic acids, and the predominating

acid in each example contains different RI and R2

groups.

Monohydroxy organic compounds useful in the preparation

of the dihydrocarbylphosphorodithioic acids

and salts useful in the present invention include alcohols,

phenol and alkyl phenols including their substituted

derivatives, e.g., nitro-, halo-, alkoxy-, hydroxy-, 55

carboxy-, etc. Suitable alcohols include, for example,

ethanol, n-propanol, isopropanol, n-butanol, 2-butanol,

2-methyl-propanol, n-pentanol, 2-pentanol, 3-pentanol,

2-methylbutanol, 3-methyl-2-pentanol, n-hexanol, 2hexanol,

3-hexanol, 4-methyl-2-pentanol, 2-methyl-3- 60

pentanol, cyclohexanol, chlorocylohexanol, methylcyclohexanol,

heptanol, 2-ethylhexanol, n-octanol,

nononanol, dodecanol, etc. The phenols suitable for the

purposes of the invention include alkyl phenols and

substituted phenols· such as phenol, chlorophenol, bro- 65

mophenol, nitrophenol, methoxyphenol, cresol, propylphenol,

heptylphenol, octylphenol, decyl phenol, dodecyl

phenol, and commercially available mixtures of

EXAMPLE 13

A mixture of 78.7 parts (1.1 equivalents) of cuprous

oxide and 112 parts of mineral oil is prepared, and 384

parts (1 equivalent) of the phosphorodithioic acid prepared

as in Example 4 are added over a period of 2

hours while raising the temperature gradually to about

55° C. Upon completion of the addition of the acid, the

reaction mixture is maintained at about 50° C. for about

3 hours. A vacuum then is applied while raising the

EXAMPLE 12

A mixture of 541 parts (13.3 equivalents) of zinc oxide,

14.4 parts (0.24 equivalent) of acetic acid and 1228

parts of mineral oil is prepared, and a vacuum is applied

while raising the temperature to about 70° C. The phosphorodithioic

acid prepared in Example 4 (4512 parts,

12 equivalents) is added over a period of about 5 hours

while maintaining the temperature at 68°_72° C. Water

is removed as it forms in the reaction, and the temperature

is maintained at 68°-72°C. for 2 hours after the

addition of phosphorodithioic acid is complete. To insure

complete removal of water, vacuum is adjusted to

about 10 mm., and the temperature is raised to about

105° C. and maintained for 2 hours. The residue is filtered,

and the filtrate is the desired product containing

6.26% phosphorus (theory, 6.09) and 6.86% zinc (theory,

6.38).

EXAMPLE 9

A mixture of 146 parts (2.5 equivalents) ofammonium

hydroxide and 40 parts of water is prepared. Beginning

at room temperature, there is added 581.4 parts (2

equivalents) of the phosphorodithioic acid prepared in

Example lover a period of 2.5 hours. The reaction is

exothermic to 40° C., and after all of the phosphorodithioic

acid is added, the reaction mixture is maintained

at 50° C. for 2 hours. An additional 59.4 parts (0.2 equivalents)

of the phosphorodithioic acid are added and the

mixture is maintained at about 50° C. for 15 hours,

cooled and filtered. The filtrate is the desired ammonium

salt which is a clear liquid.

EXAMPLE 10

To 129 parts of ammonium hydroxide (2.3 equivalents)

there is added 644.4 parts (2.0 equivalents) of the

phosphorodithioic acid prepared in Example 2 over a

period of 2 hours. The reaction is exothermic to 40° C.

After stirring for 2 hours at this temperature, the mixture

is cooled and 5 parts of ammonium hydroxide are

added through a sub-surface inlet tube. The mixture is

stirred at 40° C. for one hour whereupon 78 parts of the

isobutylamyl alcohol mixture described in Example 1

are added. The mixture is filtered through a filter aid,

and the filtrate is the desired ammonium salt containing

15.84% sulfur (theory, 14.95). '

EXAMPLE 11

A mixture of 63 parts (1.55 equivalents) of zinc oxide,

144 parts of mineral oil and one part of acetic acid is

prepared. A vacuum is applied and 533 parts (1.3 equivalents)

of the phosphorodithioic acid prepared in Exam-

35 pIe 3 are added while heating the mixture to about 80°

C. The temperature is maintained at 80°-85° C. for

about 7 hours after the addition of the phosphorodithioic

acid is complete. The residue is filtered, and the

filtrate is the desired product containing 6.8% phosphorus.

4,879,022

EXAMPLE 7

EXAMPLE 8

A mixture of 160 parts of a 50% aqueous solution of

sodium hydroxide, 40 parts of water and 200 parts ofthe 55

alcohol mixture of Example 5 is prepared, and 626 parts

of the phosphorodithioic acid of Example 2 are added

dropwise over a period of 1.5 hours. The reaction is

exothermic to 55° C., and after all of the phosphorodithioic

acid is added, the temperature of the reaction 60

mixture is increased to 65° C. and maintained at this

temperature for 2 hours. An additional 9 parts of the

50% aqueous sodium hydroxide solution are added, and

the mixture is maintained for an additional 2 hours at

55°_65° C. The mixture is fIltered through a filter aid, 65

and the fIltrate is the desired product as a 25% solution

in the alcohol mixture. The product contains 12.92%

sulfur (theory, 12.37).

reaction mixture is fIltered through a filter aid, and the

filtrate is the desired phosphorodithioic acid.

EXAMPLE 5

A mixture of 400 parts of 50% aqueous sodium hy- 5

droxide (5.7 equivalents) and 1137 parts of water is

prepared, and a mixture of 90 parts (1.1 equivalents) of

a 60/40 mixture of isobutyl alcohol/primary amyl alcohol

mixture and 1424 parts (5 equivalents) of the phosphorodithioic

acid of Example 1 is added dropwise 10

while maintaining the reaction temperature at about

40·-45· C. over a period of 4 hours. After the addition

is completed, the mixture is stirred for 45 minutes, and

an additional 56 parts of the 50% aqueous sodium hy- 15

droxide solution are added with stirring. The color of

the mixture changes from dark green to yellow, and 287

parts of water is added with stirring. The mixture, after

cooling, is filtered through a filter aid, and the filtrate is

the desired sodium salt containing 10.5% sulfur (theory, 20

9.43) and 3.52% sodium (theory, 3.86).

EXAMPLE 6

A mixture of 176 parts of a 50% aqueous solution of

sodium hydroxide, 189 parts of the alcohol mixture of 25

Example 1 and 40 parts of water is prepared, and 581.4

parts of the phosphorodithioic acid of Example 1 are

added over a period of 2 hours while maintaining the

temperature of the mixture at less than 50· C. After the

addition is completed, the mixture is maintained at 30

50·_55· C. for 2 hours and filtered. The filtrate is the

desired product containing 12.95% sulfur (theory,

12.98).

A mixture of 448 parts of zinc oxide (11 equivalents)

and 467 parts of the alcohol mixture of Example 1 is

prepared, and 3030 parts (10.5 equivalents) of the phosphorodithioic

acid of Example 1 are added at a rate to

maintain the reaction temperature at about 45·-50· C. 40

The addition is completed in 3.5 hours whereupon the

temperature of the mixture is raised to 75° C. for 45

minutes. After cooling to about 50° C., an additional 61

parts of zinc oxide (1.5 equivalents) are added, and this

mixture is heated to 75° C. for 2.5 hours. After cooling 45

to ambient temperature, the mixture is stripped to 124°

C. at 12 mm. pressure. The residue is filtered twice

through a fIlter aid, and the filtrate is the desired zinc

salt containing 22.2% sulfur (theory, 22.0), 10.4% phos- 50

phorus (theory, 10.6) and 10.6% zinc (theory, 11.1).

4,879,022

temperature to about 80· C. The residue is filtered, and

the filtrate is the desired cuprous salt which is a clear

liquid containing 12% sulfur (theory, 11.5) and 12.0%

copper (theory, 11.4).

The amount of phosphorodithioic acid or salt thereof 5

included in the slurry to be used in the flotation process

is an amount which is effective in promoting the froth

flotation process and providing improved separation of

the desired mineral values. The amount of collector

included in the slurry will depend upon a number of 10

factors including the nature and type of ore, size of ore

particles, etc. In general, from about 0.01 to about 1

pound of collector (B-2) is used per ton of ore.

The slurry prepared in step (B) also may optionally

contain (B-4) a water-soluble inorganic base. The inclu- 15

sion of a base is well known in the art for providing

desirable pH values. By controlling and modifying the

pH of the pulp slurry to levels of 8.0 and above, and

more generally above about 11 prior to conditioning,

and the pH of the slurry during the conditioning step to 20

levels of about 6.0 to 7.0 through the addition of a base,

the performance of the sulfide collectors is improved.

The alkali and alkaline earth metal oxides and hydroxides

are useful inorganic bases. Lime is a particularly

useful base. In the process of the present invention, it 25

has been discovered that the addition of a base to the

ore or slurry containing the collectors of this invention

results in a significant increase in the copper assay of the

cleaner concentrates.

The slurry (B) used in the process of the present 30

invention comprises the ground ore, at least one collector

(B-2), water and optionally a water-soluble inorganic

base. The slurries used in this invention will contain

from about 20% to about 50% by weight of solids,

and more generally from about 30% to 40% solids. 35

Such slurries can be prepared by mixing all the above

ingredients. Alternatively, the collector and inorganic

base can be premixed with the ore either as the ore is

being ground or after the ore has been ground to the

desired particle size. Thus, in one embodiment, the 40

ground pulp is prepared by grinding the ore in the presence

of collector (B-2) and inorganic base (B-4) and this

ground pulp is thereafter diluted with water to form the

slurry. The amount of inorganic base included in the

ground ore and/or the slurry prepared from the ore is 45

an amount which is sufficient to provide the desired pH

to the slurry in the subsequent conditioning step. This

amount may be varied by one skilled in the art depending

on particular preferences.

After the ore slurry has been prepared in accordance 50

with any of the embodiments described above, the

slurry is conditioned with sulfur dioxide under aeration

at a pH of from about 5.5 to about 7.5. The conditioning

medium is an aqueous solution formed by dissolving

sulfur dioxide in water forming sulfurous acid (H2S03). 55

It has been found that when the ore slurry is conditioned

with sulfurous acid and aerated, the S02 increases

the flotation rate of copper minerals, and depresses

the undesired gangue and undesirable minerals

such as iron resulting in the recovery in subsequent 60

treatment stages of a product that represents a surprising

high recovery of copper values and a surprising low

retention of iron. The amount of sulfur dioxide added to

the slurry in the conditioning step can be varied over a

wide range, and the precise amounts can be useful for a 65

particular ore or flotation process can be readily determined

by one skilled in the art. In general, the amount

of sulfur dioxide utilized in the conditioning step is

within the range of from about 1 to about 10 pounds of

sulfur dioxide per ton of ground ore. It has now been

discovered that an important factor in the conditioning

step of the process of the invention is the pH of the

slurry. The pH of the conditioned slurry should be

maintained between about 5.5 and about 7.5, more preferably

between about 6.0 to about 7.0. A pH of about

6.5 to about 7.0 is particularly preferred for the conditioned

slurry.

Conditioning of the slurry is achieved by agitating

the pulp contained in a conditioning tank such as by

vigorous aeration and optionally, with a suitable agitator

such as a motor-driven impeller, to provide good

solid-liquid contact between the finely divided ore and

the sulfurous acid. The pulp is conditioned sufficiently

long to maximize depression of the undesirable minerals

and gangue while maximizing activation of the desired

minerals such as copper minerals. Thus, conditioning

time will vary from ore to ore, but it has been found for

the ores tested that conditioning times of between about

1 to 10 minutes and more generally from about 3 to 7

minutes provide adequate depression of the undesirable

minerals and gangue.

One of the advantages of the conditioning step is that

it allows recovery of a concentrate having very satisfactory

copper content without requiring the introduction

of lime, cyanide or other conditioning agents to the

flotation circuit, although as mentioned above, the introduction

of some lime improves the results obtained.

Omitting these other conditioning agents, or reducing

the amounts of lime or other conditioning agents offers

relieffor both the additional costs and the environmental

and safety factors presented by these agents. However,

as noted below, certain advantages are obtained

when small amounts of such agents are utilized in the

flotation steps.

Following the conditioning step, the slurry is subjected

to a copper rougher flotation stage to recover

most of the copper values in the froth (concentrate)

while rejecting significant quantities of undesirable minerals

and gangue in the underflow. The flotation stage

of the flotation system, as schematically illustrated in

the figure, comprises at least one roughing stage

wherein a rougher concentrate is recovered, and one or

more cleaning stages wherein the rougher concentrate

is cleaned and upgraded. Tailing products from each of

the stages can be routed to other stages for additional

mineral recovery.

Flotation of copper is effected in the copper rougher

stage at a slightly acidic pulp pH which is generally

between about 6.0 and 7.0, the pH being governed by

the quantity of sulfur dioxide used during the conditioning

and aeration as well as the quantity of any inorganic

base included in the slurry.

The copper rougher flotation stage will contain at

least one frother, and the amount of frother added will

be dependent upon the desired froth characteristics

which can be selected with ease by one skilled in the art.

A typical range of frother addition is from about 0.04 to

about 0.1 pound of frother per ton of dry ore.

An essential ingredient of the slurry contained in the

copper rougher stage is one or more of the collectors

(B-2) described above. In one embodiment, the collector

is included in the slurry in step (B), and additional

collector may be added during the flotation steps including

the rougher stage as well as the cleaner stage. In

addition to the phosphorodithioic acids and salts, other

types of collectors normally used in the flotation of

4,879,022

with the phosphorodithioic acid and acid salts of this

invention include: sodium isopropyl xanthate, isopropyl

ethyl thionocarbamate, N-ethyl O-isopropyl thionocarbamate,

N-ethoxycarbonyl N'-isopropylthiourea, ethyl

isopropyl thionocarbamate, etc

In the rougher flotation step, the pulp is frothed for a

period of time which maximizes copper recovery. The

precise length of time is determined by the nature and

size of the ore as well as other factors, and the time

necessary for each individual ore can be readily determined

by one skilled in the art. Typically, the froth

flotation step is conducted for a period of from 2 to

about 20 minutes and more generally from a period of

about 5 to about 15 minutes. As the flotation step proceeds,

small amounts of collectors may be added periodically

to improve the flotation of the desired mineral

values. The collector added periodically to the rougher

concentrate may be additional amounts of the phosphorodithioic

acid or salt included in the slurry and/or

auxiliary collectors such as those mentioned above. In

one preferred embodiment, the collectors present during

the froth flotation comprise a mixture of one or

more of the phosphorodithioic acid salts of the invention

with one or more xanthate or thionocarbamate.

When the froth flotation has been conducted for the

desired period of time, the copper rougher concentrate

is collected, and the copper rougher tailing product is

removed and may be subjected to further purification.

The recovered copper rougher concentrate is processed

further to improve the copper grade and reduce

the impurities within the concentrate. One or more

cleaner flotation stages can be employed to improve the

copper grade to a very satisfactory level without un-

35 duly reducing the overall copper recovery of the system.

Generally, two cleaner flotation stages have been

found to provide satisfactory results.

Prior to cleaning, however, the copper rougher concentrate

is finely reground to reduce the particle size to

a desirable level. In one embodiment, the particle size is

reduced so that 60% is -400 mesh. The entire copper

rougher concentrate can be comminuted to the required

particle size or the rougher concentrate can be classified

and only the oversized materials comminuted to the

required particle size. The copper rougher concentrate

can be classified by well-known means such as hydrocyclones.

The particles larger than desired are reground to

the proper size and are recombined with the remaining

fraction.

The reground copper rougher concentrate then is

cleaned in a conventional way by forming an aqueous

slurry of the reground copper rougher concentrate in

water. One or more frothers and one or more collectors

are added to the slurry which is then subjected to a

froth flotation. The collector utilized in this cleaner

stage may be one or more of the phosphorodithioic acid

or acid salts described above as (B-2) and/or any of the

auxiliary collectors described above. In some applications,

the addition of collector and a frother to the

cleaning stage may not be necessary if sufficient quantities

of the reagents have been carried along with the

concentrate from the preceding copper rougher flotation.

Small amounts of S02 also can be added to the

copper cleaner stages. The duration of the first copper

cleaner flotation is a period of from about 5 to about 20

minutes, and more generally for about 8 to about 15

minutes. At the end of the cleaning stage, the froth

containing the copper cleaner concentrate is recovered

ROC(S)NHR'

wherein R and R' are alkyl groups. U.S. Pat. Nos.

2,691,635 and 3,907,854 describe processes for preparing

dialkylthionocarbamates as represented by the

above formula. These two patents are incorporated by 50

reference herein for their disclosures of the methods of

preparing suitable auxiliary collectors useful in this

invention.

Hydrocarboxycarbonyl thionocarbamate compounds

also have been reported as useful collectors for benefici- 55

ating sulfide ores. The hydrocarboxycarbonyl

thionocarbamate compounds are represented by the

formula

R10C(O)N(H)C(S)OR2 60

wheren R1and R2are each independently selected from

saturated and unsaturated hydrocarbyl groups, alkyl

polyether groups and aromatic groups. The preparation

of these hydrocarboxycarbonyl thionocarbamic compounds

and their use as collectors is described in U.S. 65

Pat. No. 4,584,097, the disclosure of which is hereby

incorporated by reference. Specific examples of auxiliary

collectors which may be utilized in combination

wherein R is an alkyl group containing from I to 6

carbon atoms and M is a dissociating cation such as

sodium or potassium. Examples of such xanthates in- 40

clude potassium amyl xanthate, sodium amyl xanthate,

etc.

The thionocarbamates useful as auxiliary collectors

include the dialkylthionocarbamates represented by the

formula 45

sulfide ores can be used in combination with the phosphorodithioic

acid or esters. The use of such auxiliary

collectors in combination with the collectors (B-2) of

this invention often results in improved and superior

recovery of more concentrated copper values. These 5

auxiliary collectors also may be added either to the

rougher stage or the cleaning stage, or both.

A wide variety of frothing agents have been used

successfully in the flotation of minerals from base metal

sulfide ores, and any of the known frothing agents can 10

be used in the process of the present invention. By way

of illustration, such floating agents as straight or

branched chain low molecular weight hydrocarbon

alcohols such as C6-8 alkanols, 2-ethylhexanol and

4-methyl-2pentanol (also known as methylisobutylcar- 15

binol, MIBC) may be employed as well as pine oils,

cresylic acid, polyglycol or monoethers of polyglycols

and alcohol ethoxylates.

As noted above, the froth flotation step can be improved

by the inclusion of auxiliary collectors in addi- 20

tion to the phosphorodithioic acids or salts. Any of the

known collectors can be utilized in combination with

the collectors of this invention in the rougher stage

and/or the cleaning stages of the invention. The most

common collectors are hydrocarbon compounds which 25

contain anionic or cationic polar groups. Examples

include the fatty acids, the fatty acid soaps, xanthates,

xanthate esters, thionocarbamates, dithiocarbamates,

fatty sulfates, fatty sulfonates, mercaptans, thioureas

and dialkyldithiophosphinates. The xanthates and 30

thionocarbamates are particularly useful auxiliary collectors.

One group of xanthate collectors which has been

utilized in froth flotation processes may be represented

by the formula

R-O-C(S)SM

4,879,022

13 14

and the underflow which contains the copper cleaner mary amyl alcohols. The collector is added to the pritailings

is removed. In one preferred embodiment, the mary grind, and in some instances, additional amounts

copper cleaner concentrate recovered in this manner is are added in the various process stages. In general, the

subjected to a second cleaning stage and which the ground pulp is diluted to 33% solids and conditioned

requirements for collector and frother, as well as the 5 with sulfur dioxide added as sulfurous acid for the times

length of time during which the flotation is carried out and at the pH indicated in the following tables. Typito

obtain a highly satisfactory copper content and re- cally, 6 to 7 pounds of sulfur dioxide per ton of ore are

covery can be readily determined by one skilled in the added in Examples II-IV. A rougher concentrate is

art. floated for g minutes with additional collector added

When the process of the present invention is carried 10 after 1.5, 3 and 6 minutes of flotation. Potassium amyl

out on copper sulfide ores, and in particular, Cerro, xanthate also is added as an auxiliary collector in the

Colorado mine copper sulfide ores, cleaned copper rougher stages of Examples II-IV to improve on the

concentrates are found to contain high concentrations recovery of copper middlings. The rougher concentrate

of copper with improved recoveries. recovered is ground to 60 weight percent passing 400

The following examples illustrate the process of the 15 mesh and cleaned twice.

present invention. Unless otherwise indicated in the In some of the examples, lime is added to the primary

examples and elsewhere in the specification and claims, grind to provide pH control in the subsequent sulfur

all parts and percentages are by weight, and tempera- dioxide conditioning steps, and improved results are

tures are in degrees centigrade. Also in the following obtained. In all of the examples, the frothing agent is

examples, the amounts of reagents added are expressed 20 MIBC (4-methyl-2-pentanol), also known as methyl

as "pounds per ton of dry ore" by which is meant isobutyl carbinol.

pounds of reagent per ton of fresh dry ore which is The reagent balance for four experiments conducted

ground, slurried and fed to the froth flotation system. in accordance with the process of the invention to-

The ore used in the following examples is Rio Tinto gether with a summary of the times and pulp pH of the

Cerro, Colorado mill ore (primarily chalcopyrite) as- 25 various steps are summarized in the following Tables

saying an average of about 0.54% copper, 12.5% iron I-IV. In general, Example II differs from Example I in

and 0.27% zinc. The ore is crushed to pass 10 mesh, and that lime is added to the primary grind. Example III is

ground to 50% passing 200 mesh. similar to Example II except that the collector dosage is

In the following examples, the collector utilized is the doubled. Example IV differs from Example III in that

sodium salt of the dithiophosphoric acid of Example 6 30 the regrind time is doubled.

TABLE I

Reagent Balance - Example I

Reagents Added (Ib/ton)

Stage

K-Amyl

Ca(OHh Collector Xanthate S02 MIBC

___T:.;i",m",e!...M=in'----__ Pulp

Grind Condo Froth pH

Primary Grind

Condition/Aeration

Rougher

(I)

(2)

(3)

Regrind

First Cleaner

(I)

(2)

Second Cleaner

Total

O.oI5

2.32 O.oI5

0.64 O.oI5

0.26 O.oI5

0.06 O.oro

O.oro

3.28 0.080

2.00

-lQL

O.oro

0.040

0.020

3.08 0.090

1.5

5.53

4.96

5.8

6.1

6.26

6.20

prepared from a mixture of isobutyl alcohol and pri-

TABLE II

Reagent Balance - Example II

Reagents Added Qb/ton)

K-Amyl Time. Min Pulp

Stage Ca(OHh Collector Xanthate S02 MIBC Grind Condo Froth pH

Primary Grind 7.45 0.015 11.6

Condition/Aeration O.oI5 6.5 4

Rougher

(I) 1.5 6

(2) O.ro 0.015 1.5

(3) 0.40 0.015 O.oro O.oI 1.5

(4) 0.015 0.015 O.or 3 6.7

Regrind 4

First Cleaner

(I) O.oro ..!!1.- Qm... 2 6.7

(2) 3 6.8

Second Cleaner 3 6.9

Total 7.95 0.085 0.025 6.7 0.05

4,879,022

15 16

TABLE III

Reagent Balance - Example III

Reagents Added (Ib/ton)

K-Amyl Time, Min Pulp

Stage Ca(OHh Collector Xanthate S02 MIBC Grind Condo Froth pH

Primary Grind 7.45 0.030 11.20

Condition/Aeration 0.030 6.0 4 6.8

Rougher

(I) I.S

(2) 0.030 1.5 6.7

(3) 0.030 0.010 0.01 3 6.7

(4) 0.030 0.010 0.01 3 6.7

Regrind 4

First Cleaner

(I) 0.020 0.12 0.03 2 6.7

(2) 0.020 3 6.9

Second Cleaner ~ 3 6.8

Total 7.45 0.190 0.020 6.12 0.07

TABLE IV

Reagent Balance - Example IV

Reagents Added (Ib/ton)

K-Amyl Time, Min Pulp

Stage Ca(OHh Collector Xanthate S02 MICB Grind Condo Froth pH

Primary Grind 7.45 0.030 11.30

Condition/Aeration 0.030 6.0 4 6.8

Rougher

(I) 1.5 6.8

(2) 0.030 1.5 6.8

(3) 0.030 0.010 om 3 6.7

(4) 0.030 om5 om 3 6.7

Regrind

First Cleaner

(I) 0.030 0.12 0.Q3 2 6.7

(2) om5 ~ 3 6.9

Second Cleaner 3 6.9

Total 7.45 0.195 0.025 6.12 0.07

EXAMPLE VI

In each of the above examples, the rougher concen- 40

trate and second cleaner concentrate were assayed for

percent copper and percent copper distribution. A summary

of the flotation test results is found in the following

Table V.

The general procedure of Example II is repeated

except that the sodium salt of Example 6 is replaced by

TABLE V 45 an equivalent amount of the ammonium salt of Example --------.:..=------- 9.

Summary of Flotation Test Results

Second

Calc. Rougher Concentrate Cleaner Concentrate

Ex- Head Wt. Assay % Cu WI. Assay %Cu

ample (% Cu) (%) %Cu Distr. (%) %Cu Distr. 50

I 0.507 10.69 4.02 84.7 2.53 11.8 58.9

II 0.545 7.98 5.49 84.9 1.49 23.0 62.9

III 0.526 9.72 4.82 89.0 1.62 20.7 63.7

IV 0.586 10.68 4.96 90.5 1.80 22.7 69.8

EXAMPLE VII

The general procedure of Example II is repeated

except that the sodium salt of Example 6 is replaced by

an equivalent amount of the copper salt of Example 13.

While the invention has been explained in relation to

its preferred embodiments, it is to be understood that

various modifications thereof will become apparent to

55 those skilled in the art upon reading the specification.

Therefore, it is to be understood that the invention

disclosed herein is intended to cover such modifications

as fall within the scope of the appended claims,

We claim:

1. A process of effecting the concentration of a metal

value selected from the group consisting of zinc, copper

and lead, from a sulfide ore containing said metal value

and at least one of pyrite or pyrrhotite, with selective

depression of said at least one of pyrite or pyrrhotite

65 comprising the steps of:

(A) grinding the ore to an appropriate size range;

(B) preparing a slurry comprising

(B-l) said ground ore;

EXAMPLE V

The general procedure of Example II is repeated

except that the sodium salt of Example 6 is replaced by

an equivalent amount of the zinc salt of Example 7.

As can be seen from' the results of the four examples,

the sodium dithiophosphoric acid salt is an effective

copper collector in the flotation system of the invention.

Improved results are obtained when lime is added

to the primary grind (Examples II-IV), and further 60

improvement in the copper selectivity is obtained when

the collector dosage is increased and the regrind time is

increased.

(1)

[

RiO ] ~P(S)S- xn+

R20

wherein R1 is an alkyl group containing from 2 to

about 12 carbon atoms and R2 is an aryl group

containing from 6 to about 18 carbon atoms, n is an

integer equal to the valence of X and xn+ is a

dissociating cation; and

pH of about 6.0 to about 7.0 in an amount sufficient to

depress said at least one of pyrite or pyrrhotite;

(D) subjecting the conditioned slurry to froth flotation

to produce a froth containing copper rougher concentrate

and a resultant slurry containing said at least

one of pyrite or pyrrhotite which was depressed during

the flotation;

(E) collecting said froth; and

(F) recovering the copper rougher concentrate.

(G) regrinding the recovered copper rougher concentrate;

(H) subjecting the reground concentrate to at least one

cleaning froth flotation process to form a copper

cleaner concentrate; and

(I) recovering a copper cleaner concentrate.

19. The process of claim 18 wherein the slurry of (B)

also contains a water-soluble inorganic base (B-4).

20. The process ofclaim 19 wherein the ore Is ground

in step (A) in the presence of said collector (B-2) and

inorganic base (B-4).

21. The process of claim 19 wherein the slurry of (B)

is prepared by diluting a ground pulp with water, said

ground pulp being prepared by grinding the ore in the

presence of said collector (B-2) and said inorganic base

(B-4).

22. The process of claim 18 wherein X is an ammonium

group, an alkali metal or a Group II metal.

23. The process of claim 22 wherein X is an alkali

metal.

24. The process of claim 23 wherein X is sodium.

25. The process of claim 18 wherein from about 0.01

to about 1.0 pound of said collector (B-2) is used per ton

of ore.

26. The process of claim 19 wherein the inorganic

base is an alkali or alkaline earth metal oxide, hydroxide

or mixtures thereof.

27. The process of claim 26 wherein the inorganic

base is calcium hydroxide.

28. The process of claim 18 wherein the slurry is

conditioned in step (C) at a pH of about 6.5 to about 7.0.

17. The process of claim 1 wherein at least one xanthate

or dithionocarbamate metal value collector is

added to the conditioned slurry during the froth flotation

step (D).

18. A process for the benefication of an ore containing

copper sulfide minerals and at least one of pyrite or

pyrrhotite with selective depression of said at least one

of pyrite or pyrrhotite comprising the steps of:

(A) grinding the ore to an appropriate size range;

10 (B) preparing a slurry comprising

(B-1) said ground ore;

(B-2) at least one copper collector which is a waterdispersible

or soluble dihydrocarbyldithiophosphoric

acid or salt having the formula .

(1) 5

4,879,022

(B-2) at least one metal value collector which is a

water-dispersible or soluble dihydrocarbyl dithiophosphoric

acid or salt having the formula

wherein R1 is an alkyl group contaiIiing from 2 to

about 12 carbon atoms and R2 is an aryl group

containing from 6 to about 18 carbon atoms, n is an

integer equal to the valence of X and xn+ is a 15

dissociating cation; and

pH of about 6.0 to about 7.0 in an amount sufficient to

depress said at least one of pyrite or pyrrhotite;

(D) subjecting the conditioned slurry to froth flotation 20

to produce a froth containing a metal value rougher

concentrate and a resultant slurry containing said at

least one of pyrite or pyrrhotite which was depressed

during the flotation;

(E) collecting said froth; and 25

(F) recovering the metal value rougher concentrate

containing the metal values.

2. The process of claim 1 wherein the metal value is

copper.

3. The process of claim 1 wherein the metal value is 30

lead or zinc.

4. The process of claim 1 wherein the ore is ground in

step (A) in the presence of the collector (B-2).

5. The process of claim 1 wherein the slurry of (B) is

prepared by diluting a ground pulp with water, said 35

ground pulp being prepared by grinding the ore in the

presence of the collector (B-2), and an inorganic base

(B-4).

6. The process of claim 1 wherein X is hydrogen, an

ammonium group, an alkali metal or a Group II metal. 40

7. The process of claim 6 wherein X is an alkali or

alkaline earth metal.

8. The process of claim 6 wherein the dissociating

cation is an alkali metal cation.

9. The process of claim 1 wherein the slurry (B) con- 45

tains a water-soluble inorganic base (B-4).

10. The process of claim 9 wherein the inorganic base

is an alkali or alkaline earth metal oxide, hydroxide, or

mixtures thereof.

11. The process of claim 9 wherein the inorganic base 50

is calcium hydroxide.

12. The process of claim 1 wherein the slurry is conditioned

in step (C) at a pH of from about 6.5 to about

7.0.

13. The process of claim 1 wherein the froth flotation 55

is effected in step (0) at a pH of from about 6.0 to about

7.0.

14. The process of claim 1 wherein additional mineral

value collector (B-2) is added to the conditioned slurry

during froth flotation step (D). 60

15. The process of claim 1 wherein the metal value

rougher concentrate recovered in step (F) is

(G) re-ground; and

(H) subjected to at least one cleaning froth flotation

process. 65

16. The process of claim 15 wherein additional collector

(B-2) is introduced in the cleaning froth flotation

process (H).

4,879,022

29. The process of claim 18 wherein the froth flotation

is effected in step (D) at a pH of from about 6.0 to

about 7.0.

30. The process of claim 18 wherein additional copper

collector (B-2) is added to the conditioned slurry

during froth flotation in step (D).

31. The process of claim 28 wherein additional copper

collector (B-2) is added in the cleaning froth flotation

process (H).

32. The process of claim 18 wherein at least one xanthate

or dithionocarbamate copper collector is added to

the conditioned slurry during the froth flotation.

33. The process of claim 18 wherein the slurry is

5 conditioned in step (C) by addition of from about 1 to

about 10 pounds of 502 per ton of ground ore.

34. The process of claim 18 wherein said collector

(B-2) comprises a mixture of dihydrocarbyldithiophosphoric

acids or salts of Formula I and also containing

10 minor amounts of acids or salts wherein Rl and R2 are

the same alkyl group.

* * * * *