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; .
(C) conditioning the slurry with S02 under aeration at a
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
l"
PRIMARY
GRIND
SLURRY H2O
PREPARATION COLLECTOR
CONDITIONER S02-
I
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;
(C) conditioning the slurry with S02 under aeration at a
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
2
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.
1
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
3
(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
4
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
n
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
6
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.
5
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
8
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).
7
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
9
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
10
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
12
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
11
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
0.020
3.08 0.090
4
4
1.5
2
2
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
n
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
(C) conditioning the slurry with S02 under aeration at a
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.
18
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
17
(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
(C) conditioning the slurry with S02 under aeration at a
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
19
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).
20
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.
* * * * *
15
20
25
30
35
40
45
50
55
60
65