Note: Descriptions are shown in the official language in which they were submitted.
201~882
DEPRESSION OF THE FLOTATION OF SILICA
OR SILICEOUS GANGUE IN MINERAL FLOTATION
Thi~ invention is related to the recovery of
minerals by froth flotation.
Flotation is a process of treating a mixture of
finely divided mineral solids, e.g., a pulverulent ore,
suspended in a liquid whereby a portion of the solids is
separated from other finely divided mineral solids,
e.g., silica, siliceous gangue, clays and other like
materials present in the ore, by introducing a gas (or
providing a gas in situ) in the liquid to produce a frothy
mass containing certain of the solids on the top of the
liquid, and leaving suspended (unfrothed) other solid
components of the ore. Flotation is based on the
principle that introducing a gas into a liquid
containing solid particles of different materials
suspended therein causes adherence of some gas to
certain suspended solids and not to others and makes the
particles having the gas thus adhered thereto lighter
than the liquid. Accordingly, these particles rise to
the top of the liquid to form a froth.
The minerals and their associated gangue which
are treated by froth flotation generally do not possess
-
37,014B-F ~
- 2014882
--2--
sufficient hydrophobicity or hydrophilicity to allow
adequate separation. Therefore, various chemical
reagents are often employed in froth flotation to create
or enhance the properties necessary to allow separation.
Collectors are used to enhance the hydrophobicity and
thus the floatability of different mineral values. Col-
lectors must have the ability to (1) attach to the
desired mineral species to the relative exclusion of
other species present; (2) maintain the attachment in
the turbulence or shear associated with froth flotation;
and (3) render the desired mineral species sufficiently
hydrophobic to permit the required degree of separation.
A number of other chemical reagents are used in
addition to collectors. Examples of types of additional
reagents used include frothers, depressants, pH
regulators, such as lime and soda, dispersants and
various promoters and activators. Depres~ants are used
to increase or enhance the hydrophilicity of various
mineral species and thus depress their flotation.
Frothers are reagents added to flotation systems to
promote the creation of a semi-stable froth. Unlike
both depressants and collectors, frothers need not
attach or adsorb on mineral particles.
Froth flotation has been exten~ively practiced
in the mining industry since at least the early
twentieth century. A wide variety of compounds are
taught to be u~eful as collector~, frothers and other
reagents in froth flotation. For example, xanthates,
simple alkylamines, alkyl sulfates, alkyl sulfonates,
carboxylic acids and f-tty acids are generally accepted
as useful collectors. Reagents useful as frothers
include lower molecular weight alcohols such as methyl
2014882
isobutyl carbinol and glycol ethers. The specific
additives used in a particular flotation operation are
selected according to the nature of the ore, the
conditions under which the flotation will take place,
the mineral sought to be recovered and the other
additives which are to be used in combination therewith.
While a wide variety of chemical reagents are
recognized by those skilled in the art as having utility
in froth flotation, it is also recognized that the
effectiveness of known reagents varies greatly depending
on the particular ore or ores being subjected to
flotation as well as the flotation conditions. It is
further recognized that selectivity or the ability to
selectively float the desired species to the exclusion
of undesired species is a particular problem.
Minerals and their associated ores are
generally categorized as sulfides or oxides, with the
latter group including carbonates, hydroxides, sulfates
and silicate~. While a large proportion of the minerals
existing today are contained in oxide ores, the bulk of
successful froth flotation systems is directed to
sulfide ores. The flotation of oxide minerals is
recognized as being sub~tantially more difficult than
the flotation of sulfide minerals and the effectiveness
of most flotation proces~es in the recovery of oxide
ore~ is limited.
A major problem associated with the recovery of
minerals, both oxides and sulfides, is selectivity.
Some of the recognized collectors such as the carbGxylic
acids, alkyl sulfates and alkyl sulfonates discussed
above are taught to be effective collectors for oxide
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-4- 64693-4628
201~882
mineral ores. Certainly, existing collectors are known
to be useful in sulfide flotation. However, while the
use of these collectors can result in acceptable
recoveries, it is recognized that the selectivity to the
desired mineral value may not be as high as desired and,
in the case of oxide flotation, is typically quite poor.
That is, the grade or the percentage of the desired
mineral contained in the recovered mineral is
unacceptably low.
Thus, a need remains for methods of increasing
selectivity in the flotation of both sulfide and oxide
ores.
The present invention is a process for the
recovery of mineral values by froth flotation comprising
subjecting a particulate ore, which contains silica or
siliceous gangue and is in an aqueous slurry, to froth
flotation under conditions such that the minerals to be
recovered are floated wherein the flotation of the
silica or siliceous gangue is depressed by the use of an
effective amount of a hydroxy-containing compound
selected from the group comprising ethanol amine,
propanol amine, butanol amine, diethanol amine,
dipropanol amine, tripropanol amine, triethanol amine
and mixtures thereof. Additionally, the froth flotation
process of this invention utilizes collectors, frothers
and other flotation reagents known in the art.
3o
201~882
--5
By improved selectivity, it is meant that the
total amount of mineial recovered and/or the grade of
the mineral recovered is increased while the amount of
silica or siliceous gangue not recovered, i.e. remaining
in the aqueous phase, is also increased. Thus, by the
process of this invention, the ability to separate
silica and/or siliceous gangue from desirable mineral
values is enhanced. That is, the tendency of the silica
or siliceous gangue to float is depressed.
The flotation process of this invention is
useful in the recovery of various minerals, including
oxide minerals, by froth flotation.
The flotation process of this invention is
useful in the recovery of mineral values from a variety
of ores. An ore herein refers to the mineral as it is
taken out of the ground and includes the mineral-
-containing species intermixed with gangue. Gangue are
those materials which are of little or no value and need
to be separated from the mineral values. In this
invention, gangue specifically includes silica and
siliceous materials.
As is well recognized by one skilled in the
art, different typeq of collectors are effective with
different types of ores. Certain anionic collectors,
described below and useful in the present invention,
have been found to be surprisingly effective in the
flotation of oxide ores. The oxide minerals which can
be treated by the practice of this invention include
carbonates, sulfates and silicates as well as oxides.
In addition to its effectiveness in the flotation of
2~14882
--6--
oxide ores, it has also been found that the anionic
collectors in the flotation process of this invention
are also effective in the flotation of sulfide ores and
mixed oxide/qulfide ores.
Non-limiting examples of oxide ores which can
be floated using the practice of this invention
preferably include iron oxides, nickel oxides,
phosphorus oxides, copper oxides and titanium oxides.
Other types of oxygen-containing minerals which can be
0 floated using the practice of this invention include
carbonates such as calcite or dolomite and hydroxides
such as bauxite.
The process of this invention using the anionic
collectors described below is also useful in the
flotation of various sulfide ores. Non-limiting
examples of sulfide ores which can be floated by the
process of this invention include those containing
chalcopyrite, chalcocite, galena, pyrite, sphalerite and
pentlandite.
Noble metals such as gold and silver and the
platinum group metals wherein platinum group metals
comprise platinum, ruthenium, rhodium, palladium,
osmium, and iridium, can also be recovered by the
practice of this invention. For example, such metals
are sometimes found associated with oxide and/or sulfide
ores. For example, platinum is sometimes found
associated with troilite. By the practice of the
present invention, such metals can be recovered in good
yield.
--6--
2014882
--7
Non-limiting examples of oxide ores which can
be subjected to froth flotation using the process of
this invention are those including cassiterite,
hematite, cuprite, vallerite, calcite, talc, kaolin,
apatite, dolomite, bauxite, spinel, corundum, laterite,
azurite, rutile, magnetite, columbite, ilmenite,
smithsonite, anglesite, scheelite, chromite, cerussite,
pyrolusite, malachite, chrysocolla, zincite, massicot,
bixbyite, anatase, brookite, tungstite, uraninite,
gummite, brucite, manganite, psilomelane, goethite,
limonite, chrysoberyl, microlite, tantalite and
samarskite. One skilled in the art will recognize that
the froth flotation process of this invention will be
useful for the processing of additional ores including
oxide ores wherein oxide is defined to include
carbonates, hydroxides, sulfates and silicates as well
as oxides and sulfide ores.
Ores for which the process oP this invention
using anionic thiol collectors are useful include
sulfide mineral ores containing copper, zinc,
molybdenum, cobalt, nickel, lead, arsenic, silver,
chromium, gold, platinum, uranium and mixtures thereof.
Examples of metal-containing sulfide minerals which can
be concentrated by froth flotation using the compo~ition
and proce~s of thi~ invention include copper-bearing
mineral~ such as covellite (CuS), chalcocite (Cu2S),
chalcopyrite (CuFeS2), bornite (CusFeS4), vallerite
(Cu2Fe4S7 or Cu3Fe4S7), tetrahedrite (Cu3SbS2), enargite
(Cu3(A~2Sb)S4), tennantite (Cu12As4S13)~ cubanite
(Cu2SFe4Ss), brochantite (Cu4(OH)6S04), antlerite
(Cu~S04(0H)4), famatinite (Cu3(SbAs)S4), and bou.nGnite
(PbCuSbS3); lead-bearing minerals such as galena (PbS);
antimony-bearing minerals such as stibnite (Sb2S3);
--7--
2014882
--8--
zinc-bearing minerals such as sphalerite (ZnS); silver-
-bearing minerals such as stephanite (Ag5SbS4) ard
argentite ~Ag2S`; c~ mium-bearing minerals such as
daubreelite (FeSCrS3); nickel-bearing minerals such as
pentlandite [(FeNi)gS8]; molybdenum-bearing minerals such
as molybdenite (MoS2); and platinum- and palladi~m-
-bearing minerals such as cooperite [Pt(AsS)2].
Preferred metal-containing sulfide minerals include
molybdenite (MoS2), chalcopyrite (CuFeS2), chalcocite
(Cu2S), galena (PbS), sphalerite (ZnS), bornite
(CusFeS4), and pentlandite [(FeNi)gS8].
Sulfidized metal-containing oxide minerals are
minerals which are treated with a sulfidization
chemical, so as to give such minerals sulfide mineral
characteristics. The minerals so treated can then be
recovered in froth flotation using collectors which
recover sulfide minerals. Sulfidization results in
oxide minerals having sulfide mineral characteristics.
Oxide minerals are sulfidized by contact with compoundq
which react with the minerals to form a sulfur bond or
affinity. Such methods are well known in the art. Such
compounds include sodium hydrosulfide, sulfuric acid and
related sulfur-containing salts such as sodium sulfide.
Sulfidized metal-containing oxide minerals and
oxide minerals for which this proceqs utilizing the
thiol collectorq described below iq useful include oxide
minerals containing copper, aluminum, iron, titanium,
3 magneqium, chromium, tungsten, molybdenum, manganese,
tin, uranium, and mixtures thereof. Example~ of metal-
-containing minerals which may be sulfidized by froth
flotation using the thiol collectors described below
include copper-bearing minerals such as malachite
(Cu2(0H)2C03), azurite (Cu3(0H)2(C03)2), cuprite (Cu20),
2014882
g
atacamite (Cu2Cl(OH)3), tenorite (CuO), chrysocolla
(CuSiO3); aluminum-bearing minerals such as corundum;
zinc-containing minerals such a~ zincite ~,r,O) and
smithsonite (ZnC03); tung~ten-bearing minerals such aY
wolframite [(Fe2Mn)W04]; nickel-bearing minerals such as
bunsenite (NiO); molybdenum-bearing minerals such as
wulfenite (PbMoO4) and powellite (CaMoO4); iron-
-containing minerals such as hematite and magnetite;
chromium-containing minerals such as chromite
(FeOCr203); iron- and titanium-containing minerals such
as ilmenite; magnesium- and aluminum-containing miner,als
such as spinel; titanium-containing minerals such as
rutile; manganese-containing minerals such as
pyrolusite; tin-containing ores; minerals such as
cassiterite; and uranium-containing minerals such as
uraninite, pitchblende (U20s(U308)) and gummite
( uo3nH2o ) -
Other metal-containing minerals for which the
use of thiol collectors in this proce~s is useful
include gold-bearing minerals such as sylvanite
(AuAgTe2) and calaverite (AuTe); platinum- and
palladium-bearing minerals such as sperrylite (PtAs2);
and silver-bearing minerals such a~ hessite (AgTe2).
Also included are metals which occur in a metallic
state, e.g., gold, silver and copper.
In a preferred embodiment of this invention,
copper-containing sulfide minerals, nickel-containing
3 sulfide minerals, lead-containing sulfide minerals,
zinc-containing sulfide minerals or molybdenum-
containing sulfide minerals are recovered. In an even
more preferred embodiment, a copper-containing sulfide
mineral is recovered.
_g _
2014882
Ores do not always exist purely as oxide ores
or as sulfide ores. Ores occurring in nature may
comprise both sulfur-containing and oxygen-containing
minerals as well as, in some case9, noble metals.
Metals may be recovered from the oxides found in such
ores by the practice of this invention. This may be
done in a two-stage flotation where one stage comprises
conventional sulfide flotation to recover primarily
sulfide minerals and the other stage of the flotation
utilizes the process of the present invention using the
anionic collectors described below to recover primarily
the oxide minerals. Alternatively, the various types of
minerals may be recovered simultaneously by the practice
of this invention.
In addition to the flotation of ores found in
nature, the flotation process of this invention is
useful in the flotation of oxides and sulfides from
other sources. For example, the waste materials from
various processes such as heavy media separation,
magnetic separation, metal working and petroleum
processing often contain oxides and/or sulfides that may
be recovered using the flotation process of the present
invention.
A wide variety of anionic collectors are useful
in the practice of the present invention. The anionic
portion of the anionic collector is preferably derived
from carboxylic, sulfonic, sulfuric, pho~phoric or
phosphonic acids. The anionic collector is also
hydrophobic. Its hydrophobicity is derived from a
satur2ted or unsaturated hydrocarbyl or s2turated or
unsaturated substituted hydrocarbyl moiety. Examples of
suitable hydrocarbyl moieties include straight or
--10--
2U1488~
, 1
branched alkyl, arylalkyl and alkylaryl groups. Non-
-limiting examples of substituents for the hydrocarbyl
group include alkoxy, ether, amino, hydroxy and carboxy.
When the hydrocarbyl moiety is unsaturated, it is
preferably ethylenically unsaturated. It should also be
recognized that the anionic surfactant may be a mixture
of compounds.
The anionic collector may be used in acid form
or in salt form, depending on which is soluble under
conditions of use. The appropriate form of the anionic
collector will vary depending on the particular
collector used and other conditions present in the
flotation process. One skilled in the art will
recognize that some of the anionic collectors useful in
the present invention will be soluble in the acid form
under conditions of use while others will be soluble in
the salt form. For example, oleic acid is preferably
used in the acid form and saturated carboxylic acids are
preferably used in salt form. When the anionic
collectors of the present invention are used in salt
form, the counter ion may be a calcium ion, a magnesium
ion, a sodium ion, a potassium ion or an ammonium ion.
As discussed above, the choice of an appropriate counter
ion depends on the particular anionic collector used and
its solubility. It is generally preferred that the
counter ion be a sodium ion, a potassium ion or an
ammonium ion.
Non-limiting examples of suitable anionic
collectors include linolenic acid, oleic acid, lauric
acid, linoleic acid, octanoic acid, capric acid,
myristic acid, palmitic acid, stearic acid, arachidic
acid, behenic acid, 2-naphthalene sulfonic acid, sodium
201~882
-12-
lauryl sulfate, sodium stearate, dodecane sodium
sulfonic acid, hexadecyl sulfonic acid, dodecyl sodium
sulfate, dodecyl phosphate, chloride derivative of
dodecyl phosphonic acid, 2-naphthoic acid, pimelic acid,
and dodecyl benzene sulfonate and mixtures thereof.
Preferred anionic collectors include those
derived from carboxylic acids and sulfonic acids. In
the case of the anionic surfactants derived from
carboxylic acids, the unsaturated acids such as oleic
acid, linoleic acid and linolenic acids or mixtures
thereof are preferred. Examples of mixtures of these
carboxylic acids include tall oil and coconut oil.
When the anionic collector is derived from
sulfonic acids, it is preferred to use alkyl or
alkylaryl sulfonic acids. Examples of preferred species
include dodecyl benzene sulfonic acid, dodecyl sulfonic
acid, alkylated diphenyl oxide monosulfonic acid and
salts thereof.
The thiol collectors of this invention are
compounds selected from the group consisting of
thiocarbonates, thionocarbamates, thiocarbanilides,
thiophosphates, thiopho~phinates, mercaptans, xanthogen
formates, xanthic esters and mixtures thereof.
Preferred thiocarbonates are the alkyl
thiocarbonateq represented by the structural formula:
z2
R1-Z1-C-S-M+
2014882
-l3
wherein
R1 is independently a C1_20, preferablv
C2_16, more preferably C3_12 alkyl group;
z1 and z2 are independently a sulfur or oxygen
atom; and
M+ is an alkali metal cation.
The compounds represented by this formula
include the alkyl thiocarbonates (both Z1 and z2 are
oxygen), alkyl dithiocarbonates (Z1 is 0, z2 is S) and
the alkyl trithiocarbonates (both Z1 and z2 are sulfur).
Examples of preferred alkyl monothiocarbonateq
include sodium ethyl monothiocarbonate, sodium isopropyl
monothiocarbonate, sodium isobutyl monothiocarbonate,
sodium amyl monothiocarbonate, potassium ethyl
monothiocarbonate, potassium isopropyl
monothiocarbonate, potassium isobutyl monothiocarbonate
and potassium amyl monothiocarbonate. Preferred alkyl
dithiocarbonateq include potas-qium ethyl
dithiocarbonate, sodium ethyl dithiocarbonate, potassium
amyl dithiocarbonate, sodium amyl dithiocarbonate,
potassium isopropyl dithiocarbonate, qodium isopropyl
dithiocarbonate, sodium sec-butyl dithiocarbonate,
potassium sec-butyl dithiocarbonate, sodium isobutyl
dithiocarbonate, potaqsium iqobutyl dithiocarbonate, and
the like. Examples of alkyl trithiocarbonateq include
3 sodium isobutyl trithiocarbonate and potassium isobutyl
trithiocarbonate. It is often preferred to employ a
mixture of an alkyl monothiocarbonate, alkyl
dithiocarbonate and alkyl trithiocarbonate.
2014~82
-14-
Preferred thionocarbamates correspond to the
formula:
S
(R2)a-N-C-Y
(H)b
wherein
each R2 is independently a C~_10, preferably a
C1_4, more preferably a C1_3, alkyl group;
Y is -S-M+ or -oR3, wherein R3 is a C1_10~
preferably a C2_6, more preferably a C3_4,
alkyl group;
a is the integer 1 or 2; and
b is the integer 0 or 1, wherein a+b must
equal 2.
Preferred thionocarbamates include dialkyl
dithiocarbamates (a=2, b=0 and Y is S-M+) and alkyl
thionocarbamates (a=1, b=1 and Y is -oR3). Examples of
preferred dialkyl dithiocarbamates include methyl butyl
dithiocarbamate, methyl isobutyl dithiocarbamate, methyl
sec-butyl dithiocarbamate, methyl propyl
dithiocarbamate, methyl isopropyl dithiocarbamate, ethyl
butyl dithiocarbamate, ethyl isobutyl dithiocarbamate,
ethyl sec-butyl dithiocarbamate, ethyl propyl
dithiocarbamate, and ethyl i~opropyl dithiocarbamate.
3 Examples of preferred alkyl thionocarbamates include
N-methyl butyl thionocarbamate, N-methyl isobutyl
thionocarbamate, N-methyl sec-butyl thionocarbamate,
N-methyl propyl thionocarbamate, N-methyl isopropyl
thionocarbamate, N-ethyl butyl thionocarbamate, N-ethyl
isobutyl thionocarbamate, N-ethyl sec-butyl
-14-
2014882
_ -15-
thionocarbamate, N-ethyl propyl thionocarbamate, and
N-ethyl isopropyl thionocarbamate. Of the foregoing,
N-ethyl isopropyl thionocarbamate and N-ethyl isobutyl
thionocarbamate are most preferred.
Thiophosphates useful herein generally
correspond to the formula:
R40 S
~ P-Z--M+
R40
~ wherein each R4 is independently hydrogen or a Cl_1o
alkyl, preferably a C2_8 alkyl, or an aryl, preferably
an aryl group having from 6-10 carbon atoms, more
preferably cresyl; Z is oxygen or sulfur; and M is an
alkali metal cation.
Of the thiophosphates, those preferably
employed include the monoalkyl dithiophosphates (one R4
is hydrogen and the other R4 is a C1_10 alkyl and Z is
S), dialkyl dithiophosphates (both R4 are C1_l0 alkyl
and Z is S) and dialkyl monothiophosphate (both R4 are a
C1_10 alkyl and Z is 0).
Examples of preferred monoalkyl
dithiophosphates include ethyl dithiophosphate, propyl
dithiophosphate, isopropyl dithiophosphate, butyl
dithiophosphate, sec-butyl dithiophosphate, and isobutyl
dithiophosphate. Examples of dialkyl or aryl
dithiophosphates include sodium diethyl dithiophosphate,
sodium di-sec-butyl dith ophosphate, sodium diisobutyl
dithiophosphate, and sodium diisoamyl dithiophosphate.
Preferred monothiophosphates include sodium diethyl
2014882
-16-
monothiophosphate, sodium di-sec-butyl
monothiopho~ph~te, sodium diisobutyl monothiophosphate,
and sodium dii~oamyl monothiopho~phate.
Thiocarbanilides (dialkyl thioureas) are
represented by the general structural formula:
( R5-N~2C - S
wherein each R5 is individually H or a C1_6, preferably
a C1_3, hydrocarbyl.
Thiophosphinates are represented by the general
structural formula:
~S
(R6)2 -- P
\ S-M+
wherein M+ is as hereinbefore described and each R6 is
independently an alkyl or aryl group, preferably an
alkyl group having from l to 12, more preferably an
alkyl group having from l to 8 carbon atoms. Most
preferably, each R6 iq isobutyl.
Mercaptan collectors are preferably alkyl
mercaptans repreqented by the general structural
3 formula:
R7-S-H
-16-
2014882
-17-
wherein R7 is an alkyl group, preferably an alkyl group
having at least 10, more preferably from 10 to 16,
ea~bon atoms.
Xanthogen formates are represented by the
general structural formula:
R8 __O _ c~i ~ O
~S--C
OR9
wherein R8 is an alkyl group having from 1 to 7 ,
preferably from 2 to 6 carbon atoms and R9 is an alkyl
group having 1 to 6, preferably 2 to 4, more preferably
2 or 3, carbon atoms.
Xanthic esters are preferably compounds of the
general structural formula:
R10_-O -C
~ S - Rl1
wherein RlO i~ an allyl group and Rll is an alkyl group
3 having from l to 7 carbon atoms.
Preferred thiol compound~ for use a~ a
collector are the thiocarbonates, thionocarbamates and
the tniophosphates due to the surprisingly high
-17-
2014882
-18-
recoveries and selectivities towards mineral values
which can be achieved.
As will be recognized by one skilled in the
art, the thiol collectors described above are
particularly useful in the flotation of sulfide minerals
or sulfidized oxide minerals. The other anionic
collectors described above are useful in the flotation
of certain sulfide minerals, but are also surprisingly
useful in the flotation of oxide minerals.
The hydroxy-containing compounds useful in the
practice of this invention comprises compounds
containing at least one -OH moiety. This hydroxy
compound is selected to be essentially non-frothing
under the conditions of use. For purposes of this
invention, non-frothing compounds are those which have
minimal frothing action under the conditions of use. As
is well recognized by those skilled in the art, when
considering simple hydroxy-containing compounds such as
alcohols, their frother power generally increases with
the number of carbon atoms in the alcohol up to about
six or seven. When the number of carbon atoms reaches
this point, the effectiveness of the alcohol as a
frother drops. Thus, under some conditions of use,
monohydric alcohols such as octanol, nonanol, decanol,
undecanol and dodecanol may be useful as non-frothing
hydroxy-containing compounds. Laboratory scale
flotation work using relatively pure water has shown
3 that these alcohols may be non-frothing and useful in
the practice of this invention. However, under most
practical conditions of use, these alcohols demonstrate
sufficient frothing so that their use is not preferred.
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- 2014882
-19- 64693-4628
The hydroxy-containing compound i9 preferably
an slkanol amine, even more preferably a lower alkanol
amine. Non-limiting examples of lower alkanol amines
uqeful in the practice of this invention include ethanol
amine, propanol amine, butanol amine, diethanol amine,
dipropanol amine, tripropanol amine, triethanol amine
and mixtures thereof.
The alkanol amines usefuL in the practice of
this invention are available commercially. As will be
recognized by one skilled in the art, commercially
available alkanol amines will have varying degrees of
purity. For example, diethanol amine may contain
varying amounts of ethanol amine and/or triethanol
amine. Such alkanol amines are suitable in the practice
of the present invention.
The hydroxy-containing compounds may be added
directly to the float cell or may be added to the
zO grinding stage. The preferred time of addition will
vary depending on the particular ore being floated, the
other reagents present and the processing system being
used. The hydroxy-containing compounds are not
pre-mixed with the collector prior to addition to the
flotation process. They are preferably added to the
Elotation system separately from the collector. They
are also preferably added prior to the addition of the
collector. For example, the hydroxy-containing
compounds may be added to the grinding stage.
3o
The collector can be used in any concentration
which gives the desired recovery of the desired metal
values. In particular, the concentration used is
2014882
, o~O
dependent upon the particular mineral to be recovered,
the grade of the ore to be subjected to the froth
flotation proceqs and the desired quality of the mineral
to be recovered. Additional factors to be considered in
determining dosage levels include the amount of surface
area of the ore to be treated. As will be rec~gnized by
one skilled in the art, the smaller the particle size,
the greater the amount of collector reagents needed to
obtain adequate recoveries and grades.
Preferably, the concentration of the collector
is at least about 0.001 kg/metric ton, more preferably
at least about 0.005 kg/metric ton. It is also
preferred that the total concentration of the collector
is no greater than about 5.0 kg/metric ton and more
preferred that it is no greater than about 2.5 kg/metric
ton. It is more preferred that the concentration of the
collector is at least about 0.005 kg/metric ton and no
greater than about 0.100 kg/metric ton. It is generally
preferred to start at the lower concentration range and
gradually increase the concentration to obtain optimum
performance.
The concentration of the hydroxy-containing
compounds useful in this invention is preferably at
least about 0.001 kg/metric ton and no greater than
about 5.0 kg/metric ton. A more preferred concentration
is at leaYt about 0.005 kg/metric ton and no more than
about 0.500 kg/metric ton. As discussed above, it is
generally preferred to start at the lower concentration
range and gradually increase the concentration to obtain
optimum. perform nce. This is particularly important
when thiol collectors are used in the flotation of
sulfide minerals since the general trend is that
201~882
selectivity is increased at the expense of overall
reccvery.
It has been found advantageous in the recovery
of certain minerals to add the collector to the
flo~ation system in stages. By staged addition, it is
meant that a part of the total collector dose is added;
froth concentrate is collected; an additional portion of
the collector is added; and froth concentrate is again
collected. This staged addition can be repeated several
times to obtain optimum recovery and grade. The number
of stages in which the collector is added is limited
only by practical and economic constraints. Preferably,
no more than about six stages are used.
In addition to the collectors and hydroxy-
-containing compounds of this invention, other
conventional additives may be used in the flotation
process, including other collectors. Examples of such
additives include depressants and dispersants. In
addition to these additives, frothers may be and
preferably are also used. Frothers are well-known in
the art and reference thereto is made for the purposes
of this invention. Non-limiting examples of useful
frothers include C5_8 alcohols, pine oils, cresols, C1-6
alkyl ethers of polypropylene glycols, dihydroxylates of
polypropylene glycols, glycol fatty acids, soaps,
alkylaryl sulfonates and mixtures thereof.
When the anionic collectors of this invention
are used, pH is theorized to play a role in the
flotation process. The r.~ture of the ~nionic collectors
of the present invention is related to the charge
characteristics of the particular oxide mineral to be
2014882
.~
recovered. Thus, pH plays an important role in the
froth flotation proces~ of the present invention. While
not wishing to be bound by any particular theory, it i~
assumed that the anionic collector attaches to the oxide
at least in part through charge interaction with the
mineral surface. Thus ! pH conditions under which the
charge of the oxide mineral is suitable for attachment
are required in the practice of this invention.
The pH in flotation systems may be controlled
by various methods known to one skilled in the art. A
common reagent used to control pH is lime. However, in
the practice of this invention, it is preferred to use
reagents such as potassium hydroxide, sodium hydroxide
and sodium carbonate and other reagents having
monovalent cations to regulate pH. Reagents having
divalent cations such as magnesium hydroxide and calcium
hydroxide may be used, but are not preferred since their
use re~ults in the need to use larger dosages of the
collector. It should be noted that when the anionic
collector is derived from sulfonic and sulfuric acids,
the presence of divalent and/or metal cations is not as
detrimental.
The following examples are provided to illus-
trate the invention and should not be interpreted as
limiting it in any way. Unless stated otherwise, all
parts and percentages are by weight.
3o
The following examples include work involving
Hallimond tube flotation and flotation done in
laboratory scale flotation cells. It should bQ noted
that Hallimond tube flotation is a simple way to screen
collectors, but does not necessarily predict the success
2014882
A
of collectors in actual flotation. Hallimond tube
flotation does not involve the shear or agitation
present ir. actual f'otation and does not measure the
effect of frothers. Thus, while a col'ector must be
effective in a Hallimond tube flotation if it is to be
effective in actual flotation, a collector effective in
Hallimond tube flotation will not necessarily be
effective in actual flotation. It should also be noted
that experience has shown that collector dosages
required to obtain satisfactory recoveries in a
Hallimond tube are often substantially higher than those
required in a flotation cell test. Thus, the Hallimond
tube work cannot precisely predict dosages that would be
required in an actual flotation cell.
Example 1 - Hallimond Tube Flotation of
Malachite and Silica
In this example, the effect of various col-
lectors on the flotation of copper was determined using
a Hallimond tube. About 1.1 g of (1) malachite, a
copper oxide mineral having the approximate formula
Cu2C03(0H)2, or (2) silica was sized to about -60 to
+120 U.S. mesh and placed in a small bottle with about
20 ml of deionized water. The mixture was shaken 30
seconds and then the water phase containing some
~uspended fine solids or slimes was decanted. This
desliming step was repeated ~everal times.
3o
A 150-ml portion of deionized water was placed
in a 250-ml glass beaker. Next, 2.0 ml of a 0.10 molar
aolut-on o pCtaQ-Q~ um r.itrate was added as a buffer
electrolyte. The pH was adjusted to about 10.0 with the
addition of 0.10 N HCl and/or 0.10 N NaOH. Next, a
~ ?~ 2014882
rA
1.0-g portion of the deslimed mineral was added along
with deionized water to bring the total volume to about
~80 ml. The collector and hydroxy-containirg compound,
as identified in the variou~ runs reported in Table I
below, were added and allowed to condition with stirring
for 15 minutes. The pH was monitored and adjusted as
necessary.
The slurry was transferred into a Hallimond
tube designed to allow a hollow needle to be fitted at
the base of the 180-ml tube. After the addition of the
slurry to the Hallimond tube, a vacuum of 5 in (12.7 cm)
of mercury was applied to the opening of the tube for a
period of 10 minutes. This vacuum allowed air bubbles
to enter the tube through the hollow needle inserted at
the base of the tube. During flotation, the slurry was
agitated with a magnetic stirrer set at 200 revolutions
per minute (RPM).
The floated and unfloated material was filtered
out of the slurry and oven dried at 100C. Each portion
was weighed. After each test, all equipment was washed
with concentrated HCl and rinsed with 0.10 N NaOH and
deionized water before the next run.
The results obtained using the above-described
procedure and varying the identity of the collector and
hydroxy-containing compound are reported in Table I
following. The recovery of malachite and silica,
respectively, reported is that fractional portion of the
original mineral placed in the Hallimond tube that is
recovered. Thus, a recovery of 1.00 indicates that all
of the material was recovered. It should be noted that
although the recovery of copper and silica,
201~882
respectively, is reported for each run, the data is
actually collected in two separate experiments done
under identical con~itions. It should further be noted
that a low silica recovery suggests a selectivity to the
copper. The values given for copper recovery generally
are correct to +Q.05 and those for silica recovery are
generally correct to +0.03.
- 201~8B2
-26- 64693-46Z8
TABLE I
Frac- Frac-
tional tional
Dossge Cu Re- Silica
Run Collector (kQ/kQ) coverY Recovery
1~Olelc acld 0.024 0.860 0.096
2~Laurlc acld 0.024 0.786 0.154
3~Octanoic acld 0.024 0.228 0.354
4~Llnolelc acld 0.024 0.982 0.120
5~2-naphthalene O.OZ4 0.073 0.000
sulfonlc acld
6~Sodium lauryl 0.024 0.971 0.106
sulfate
7~Dodecyl sodlum 0.024 0.223 0.212
sulfonate
8~Dodecyl phosphonic 0.024 0.910 0.071
acld
9~1,2-dodecanedlol 0.024 0.255 0.210
10~1,2-dodecanediol 0.012 0.938 0.154
Olelc acld 0.012
11~Benzolc acld 0.024 0.058 0.000
~_ 12~Benzoic acld 0.01Z 0.592 0.071
Oleic acld 0.012
3o
2014882
-
-27- 64693-4628
TABLE I (Cont'd.)
Frac- Frac-
tional tional
Dosage Cu Re- Silica
RunCollector (k~/k~) covery Recovery
13~Hydroxy benzoic acld0.024 0.072 0.246
14~Hydroxy benzolc 0.012 0.732 0. 191
acid
Oleic acid 0.012
15~Trlhydroxy benzo~c 0.24 0.068 0.113
acld
_ 16~Trlhydroxy benzolc 0.012 0.816 0.089
acld
Olelc acld 0.012
17~Phenyl 0.024 0.059 0.137
18~Phenol 0.012 0.389 0.099
Oleic acid 0.012
19~Pota~slum ~alt of 0.024 0.962 0.137
dodecyl xanthate
20~C6Hg(CH2)20Cs2~ 0.024 0.170 0.165
21~Linolenlc acld 0.024 0.973 0.243
Z2~Stearlc acid 0.024 1.000 0.122
23~Palmltlc acld 0.024 1.000 0.082
24~Clycerol 0.024 0.038 0.380
25~Glycerol 0.012 0.748 0.283
Olelc acld 0.012
.
r~
A
- 2014882
-28- 64693-4628
TABLE I (Cont'd.)
Frac- Frac-
tional tional
Dosage Cu Re- Silica
RunCollector tkQ/kQ) covery Recovery
26~Ethanol amlne 0.024 0.435 0.261
27Ethanol amlne 0.012 0.963 0.105
Oleic acld 0.012
28~2-propanol amlne 0.024 0.541 0.294
10 292-propanol amlne 0.012 0.993 0.117
Olelc acld 0.012
30~Glycolic acld 0.024 0.116 0.049
31~Clycollc acld 0.012 0.904 0.047
Oleic acld 0.012
32G~ p-hydroxy propion~c acid O. 024 0.247 0.061
33~ p-hydroxy propion~c llcld 0.012 0.933 0.060
Olelc Acid 0.012
34~ Lactlc Acld 0.024 0.094 0.035
35~ Lactlc Acld 0.012 o.893 0.031
Olelc acld 0.012
36~3-hydroxy-1-propane 0.024 o.513 0.119
sulfonlc acld
37~3-hydroxy-1-propane 0.012 0.971 0.090
sulfonlc acld
Olelc acid 0.012
3o
--- 201~882
-29- 64693-4628
TABLE I (Cont'd.)
Frac- Frac-
tional tional
Dos~ge Cu Re- Sllica
RunCollector (kQ/kQ) covery Recovery
38~Propylene glycol 0.024 0.344 0.149
39(~Propylene glycol 0.012 0.967 0.077
Oleic acld 0.012
40~Propylene glycol 0.012 0.917 0.051
Laurlc acld 0.012
41~Propylene glycol 0.012 0.855 0,099
Octanolc Acld 0.012
42~Propylene glycol 0.012 o.g79 o.ol9
Linoleic acld 0.012
43~Propylene glycol 0.012 0.391 0.020
2-naphthalene 0.012
sulfonlc acid
44~Propylene glycol 0.012 O. 994 0.068
Sodium Lauryl 0.012
Sulfate
45~Propylene glycol 0.012 0.844 0.092
Dodecyl ~odium 0.012
sulfonate
46~Propylene glycol 0.012 0.998 0.088
Potassium salt of 0.012 -
dodecyl xanthate
47~Propylene glycol 0.012 0.773 0.061
C6H9 ( CH2)20CSzK 0.012
48~Propylene glycol 0.012 1.000 0.067
Linolenlc acld 0.012
49~Propylene glycol 0.012 1.000 0.099
Stearlc acld 0.012
201~882
-30- 64693-g628
TABLE I (Cont'd.)
Frac- Frac-
tional tional
Dosage Cu Re- Silica
RunCollector (kR/kR)coverYRecovery
50~Propylene glycol 0.012 1.000 0.049
Palmltlc acid 0.012
51~Propylene glycol 0.0120.818 0.043
Dodecyl benzene 0.012
sulfonic acid
10 52~Dlethanol amlne 0.0240.389 0.147
53Diethanol amlne 0.0121.000 0.071
Oleic acid 0.012
54Diethanol amine 0.0120.991 0.023
Linoleic acid 0.012
55Diethanol amine 0.0120.791 0.097
Dodecyl sodium 0.012
sulfonate
56Diethanol amine 0.0120.801 0.047
Dodecyl benzene 0.012
sulfonic acid
57~Amlno decanol 0.0240.197 0.071
58~Amino decanol 0.0120.731 0.047
Oleic acid 0.012
~ Not an embodiment of the invention
3o
- 201~882
The data in the table above indicates the broad
effectiveness of the present invention in a Hallimond
tube. ~t also indicates that the hy~rcxy-containing
compound alone generally functions poorly as a
col~ector.
Example 2 - Hallimond Tube Flotation of
Chrysocolla and Silica
The procedure outlined for Example 1 is
folIowed with the exception that chrysocolla
(Cu2H2Si205(0H)4) is used in place of malachite. In
addition, in some cases different collector~ and
hydroxy-containing compounds are used. The results
obtained are ~et out in Table II following.
3o
201~882
-32- 64693-4628
T~BLE II
Frac- Frac-
tional tional
Dosage Cu Re- Silica
RunCollector (kQ/kR) covery Recovery
1~Oleic acid 0.024 0.950 0.137
2~Dodecyl benzene 0.024 0.363 0.163
sulfonlc acld
3~Propylene glycol 0.024 0.227 0.146
4~Dlethanol amlne 0.024 O. 191 O. 151
5~Propylene glycol 0.012 0.999 0.094
Oleic acid 0.012
6~Propylene glycol 0.012 0.844 0.101
Dodecyl benzene 0.012
sulfonlc acld
7Diethanol amlne 0.012 0.986 0.096
Oleic acid 0.012
8Dlethanol amlne 0.012 0.773 O. 119
Dodecyl benzene 0.012
sulfonlc acld
Not arl embodime~t o~ the invention
The data in Table II above demon~trates the
general efEectiveness oE the present invention in the
recovery of copper from chrysocolla in Hallimond tube
Elotation within the limitations discussed relating to
Example l. These runs demonstrate that the use of tl1e
hydroxy-containing compound and anionic surfactant
results in increased copper recovery, decreased silica
recovery or both when compared to identical runs using
either component alone.
~.
201~882
~rA
Example 3 - Flotation of Mixed Copper Oxide Ore
In this example, the effect of different col-
lectorq and hydroxy-containing compounds on the
flotation of copper ore in laboratory flotation cells
was examined. Samples of copper ore from Central Africa
containing 500 g per sample were prepared. The ore
contained about 76 percent by weight malachite and the
remainder was made up of chrysocolla and chalcocite. A
500-g portion of the ore was ground with 257 g deionized
water in a rod mill at about 60 RPM for two minutes.
The resulting pulp was next deslimed. The pulp
was placed in a flotation cell. The cell was filled
with water, the slurry pH adjusted to 9.2 with sodium
carbonate and then stirred for 5 minutes. The solids in
the cell were allowed to settle for 120 secondq and then
the water phase containing finely divided solids was
decanted. This process was repeated four times. This
deslimed pulp was used in Run 8. In Runs 1-7, the
desliming steps were omitted.
The pulp was transferred to a 1500-ml Agitair
Flotation cell outfitted with an automatic paddle
removal ~ystem. The pH of the slurry was adjusted to
9.2 by the addition of sodium carbonate, if necessary.
The collectors and hydroxy-containing compounds
specified in Table III were added separately to the
slurry in the amounts specified in Table III and the
slurry was allowed to condition for one minute after the
addition of each. A polyglycol ether frother, in the
amount of 40 g per ton of dry ore, was then added and
cthe slurry allowed to condition for one additional
minute.
~ c 20148~2
r~
The flotation cell was agitated at 1150 RPM and
air introduced at a rate of 4.5 liters per minute.
Samples of the froth concentrate were collected at 1.0
and 6.0-minute intervals after the air was first intro-
duced into the cell. Samples of the tailings and con-
centrate were dried, weighed, and pulverized for analy-
sis. After being pulverized, they were dissolved with
the use of acid and the copper content determined using
a DC Plasma spectrometer. The assay data was used to
determine fractional recoveries and grades using
standard mass balance formulas.
The data obtained is shown in Table III
following.
TABLE III
Copper Recovery and Grade
Dosage
RunCollector (kg/met- 0-1 Minute1-6 Minutes Total
ric ton)
Rec Gr Rec Gr Rec Gr
1~ NaSH 0.5 0.156 0.091 0.085 0.048 0.241 0.076
C5HllOCs2K 0.2
2~Diethanol 0.2 -- -- -- -- 0.061 0.057
amine
3Diethanol 0.1 0.508 0.061 0.117 0.029 0.625 0.055
amine
Oleic acid 0.1
4~Ethanol amine 0.2 -- -- -- -- 0.044 0.058
5Ethanol amine 0.1 0.463 0.072 0.096 0.037 0.559 0.066
Oleic acid 0.1
o
00
TABLE III (Cont'd)
Copper Recovery and Grade
Dosage
RunCollector (kg/met- 0-1 Minute1-6 Minute~ Total
ric ton)
Rec Gr Rec Gr Rec Gr
6~32-propanol 0.2 -- -- -- -- 0.056 0.048
amine
7 2-propanol 0.1 0.510 0.059 0.084 0.030 0.594 0.055
amine
Oleic acid 0.1
8~Oleic acid 0.2 0.549 0.058 0.021 0.009 0.570 0.056
Not an embodiment of the invention.
oo
00
3~, 2014882
-
The data in Table III above demonstrates the
effectiveness of this invention under conditions
approximating actuai flotation conditions. Run 1, which
is not an example of the invention, approximates current
industry practice. Runs 3, 5, and 7, which are examples
of the invention, demonstrate the effectiveness of the
process of this invention in the recovery of copper.
Example 4 - Flotation of Chrysocolla Ore
A series of samples containing 500 g of ore
from Central Africa were prepared. The ore contained
greater than 90 percent chrysocolla and the remainder
comprised additional copper oxide minerals and gangue.
A 500-g sample was ground with 257 g of deionized water
in a rod mill at about 60 RPM for six minutes. The
resulting pulp was transferred to an Agitair 1500 ml
flotation cell outfitted with an automated paddle
removal system. The pH of the slurry was adjusted by
the addition of either sodium carbonate or HCl. The
natural ore pH in slurry form was 7.8. After addition
of the hydroxy-containing compounds as shown in Table
IV, the slurry was allowed to condition for one minute.
The collector was then added followed by an additional
minute of conditioning. A polyglycol ether frother was
added in an amount of 20 g per ton of dry ore followed
by an additional minute of conditioning.
3o
The float cell was agitated at 1150 RPM and air
is introduced at a rate of 4.5 liters per minute.
Samples OL the D-oth col¢ate we,-~ collected at 1.0
and 6.0 minute intervals after the air was first
2014882
. 3~
introduced. The samples of the concentrates and the
tailings were dried, weighed, pulverized for analysis
and diqqolved wlth the use of acid. The copper content
was determined by the use of DC Plasma Spectrometer.
Using the assay data, fractional recoveries and grades
were calculated using standard mass balance formulas.
The results obtained are shown in Table IV following.
3o
-- _39_ 2 01~88~ 64693-4628
~1 ' ~ ' a~
O O O O O o b~
o o
O o o o o o S
S-.
oo. ~. j`n. ! s
~ , o o o o
- ~1 . ~ . 1; il
L O o O e
Q , ~O
o C~ O ~ O C
v ~1 . . . , . , v
O O O O
.~
~1 ~ 3 1 ~r I ~
E~ o o O O av~
C
c~ OE
b~ E V In O ~
o bo ~ '`J v~ ~
a ~ .~~ ~
C C o
L ~ ~ C ~ r
~ ~ CJ- ~,
c ~ ZL~ L
t~ O
2014882
~, ~
The data in Table IV generally demonstrates the
effectiveness ~f the ~ollector composition of the
present invention. Run 2 approximates current industry
standards.
Example 5 - Flotation of Iron Oxide Ore
A series of 600-g samples of iron oxide ore
from Michigan were prepared. The ore contained a
mixture of hematite, martite, goethite and magnetite
mineral species. Each 600-g sample was ground along
with 400 g of deionized water in a rod mill at about 60
RPM for 10 minutes. The resulting pulp was transferred
to an Agitair 3000 ml flotation cell outfitted with an
automated paddle removal system. The pH of the slurry
was adjusted from a natural pH of 7.3 to a pH of 8.5
using sodium carbonate. The hydroxy-containing
compound, if used, was added and the slurry allowed to
condition for one minute. This was followed by the
addition of the collector, followed by an additional
minute of conditioning. Next, an amount of a polyglycol
ether frother equivalent to 40 g per ton of dry ore was
added followed by another minute of conditioning.
The float cell wa~ agitated at 900 RPM and air
introduced at a rate of 9.0 liters per minute. Samples
of the froth concentrate were collected at 1.0 and 6.0
minutes after the start of the air flow. Samples of the
froth concentrate and the tailings were dried, weighed
and pulverl~ed for analys s. They were then dissolved
4~
201~882
r~ 4~
in acid, and the iron content determined by the use of a
D.C. Plasma Spectrometer. Using the assay data, the
fractional recoveries and gradea were ~alculated using
standard mass balance formulas. The results are shown
in Table V following.
3o
4/
201~882
-42- 64693-4628
~ N -- ~ `D ~ C~
L I ~ J J O ~ U~ ~O
~ ~1 ~ (~ J ~ J :~' S
,_~ O O O O O O O
o
E-- o ~ Il~ O
c~: ~o o In ~ ~
O O O O O O O
td
C~
~o o ~ 0 t--
~7 . ~O :r ~1 0 U~ N U~
Ca) c~ ~ ~ 3 ~
O O O O O O O
~._
O~o c) ~ ~7 a~ ~ u~
O O ~ O O O
O O O O O O O
C
o
O ~
~ v C~ ¦ ~ ~ J ~ :~ J ~r
L C
C~
. C10 J ~ ~n t-- ~ ~
o ~ ~ O ~r ~ In ~ ~r
~:
o o o o o o O
I C
o o In u~ o o o o o o o
N ~I o o , ~ ~ , ~ ,_ ~
b~ '
c
C~ C ~, c ~, ~ ~ c . E
d ~ O ~ O ~ ~ a) O ~ ~ O
~ ~ bO bO a~ bO a~
t~ O t4 0~ 0 o ~ o a a~
.~
a
TABLE V (Cont'd.)
Iron Recovery and Grade
Dosage
Run Collector (kg/met- 0-1 Minute1-6 Minutes Total
ric ton)
Rec Gr Rec Gr Rec Gr
8'~ Diethanol 0.200 -- -- -- -- 0.141 0.458
amine
9~ Ethanol 0.200 -- -- -- -- 0.074 0.376
amine
Ethanol 0.100 0.298 0.357 0.089 0.396 0.387 0.366
i~ amine
Oleic acid 0.100
~Not an embodiment of the invention.
~Only one concentrate sample collected.
o
00
00
2014882
~A
The data in Table V above demonstrates the
effectiver.ess of the present inventiGn in obtaining good
recoveries of high grade iron.
Example 6 - Flotation of Arizona Copper Oxide Ore
A serieq of 30-g samples of -60 mesh copper ore
from Arizona were prepared. It should be noted that
this ore is very fine and, thus, very difficult to
float. The make-up of the valuable components of the
ore was about 60 percent azurite [Cu3(C03)(0H)2],
35 percent malachite [Cu2C03(0H)2], and 5 percent
chalcocite [Cu2S]. Each sample of ore was ground with
15 g of deionized water in a rod mill (2.5 inch diameter
with 0.5 inch rods)(6.35 cm dia. with 1.27 cm rods) for
240 revolutions. The resulting pulp was transferred to
a 300 ml flotation cell.
The pH of the slurry was left at natural ore pH
of 8.0 unless otherwise noted. After addition of the
hydroxy-containing compound as shown in Table VI, the
~lurry was allowed to condition for one minute. Next,
the collector was added with an additional minute of
conditioning. Next, the frother, a polyglycol ether,
was added in an amount equivalent to 0.050 g per ton of
dry ore and the slurry allowed to condition an
additional minute.
The float cell was agitated at 1800 RPM and air
introduced at a rate of 2.7 li~er~ per minute. Samples
~ 2014882
~- ~f
of the froth concentrate were collected by standard hand
paddling at 1~0 and 6.0 minutes after the start of the
introduction of air into the celi. Sampl~ of the
concentrate and the tailings were dried and analyzed as
described in the previous examples. The results
obtained are presented in Table VI following.
3o
2014813~
,
-46- 64693-4628
~I J ~ t-- ~ ll~
~,,1 ~ ~ :~ W ~ C~ ~)
Vl I o o o
_~ O O O O O O O
o
c~ t--
C~ In ~ ~ In I o o~
a~ ac ~ ~n J ~ ~ tn O ~
~ O O O O O O O
C~
u~ 0 o ~r
C~ ~ ~ 0 1 0
I o
,, 1 0 0 0 0 0
n a~
c~ In ~ O~
a~ ac N ~ I .
O O O O O
0 0 0 J ~
o o
C
O U~
C~O~ O
O ~O ~ ~J N
O O O O O
I C
~ o g 0 0 0 O O O O O
~ J ~ N ~ J t~ o J J
O bO,_~ o
C~ C ~ Cl C C C
~ ~, ~ ~-1 ~ ~ O " O
C ) O ~ a o a~ O O a~
acl ~ ~1 ~1 t (~
,
201~882
-
-47- 64693-4628
N t--
~1
~-1 0 0 ~ O O
o
E- o U~ t~ ~
CJ ~O O ~
a~ ~ ~ o ~In
a ~:
O O O OO
Id
C~
d a.),,~,1 I~ o
~ 3 0 0
O t--N
V ~OCJ I Jt_ ,_
~;
o o o
~ 3
C ~ ~ I o o I IO
C O
~~) J
J O ~ ~J N ~ ~.
d O O
t-
~ ~ ~ OCO~ O O
C ~ ~ 1 N ~ SN N
~ L N
~ C,, ~ ~ , 4, C ~
N
C~ <~ O~ O
h~
201~882
-48- 64693-4628
. ~ ~ t--
C~ o o o o
,_, o oo o
J~
o a
E~ ~ ~O
C~ O~ J ~ ~ J
Cl: O~ a J ~_,
O OO O
~_ C
C~ O
~1 , , a
. O ~
C ' O
~ _1 C~
O ~ O
V ~ O J
O O
C
O
Q ~
Q J ~I ~ al
C'3 ~ ~1 1 1 ~
C O O ~ N
._, 3
~ C
m o ~¦ I I~ ~ c ~
~: o o o C) C
~ a~
I C C ~ ~
b~) oO o o oc ~ c
V` ~ J ~~ `J ~ ~~ O Q
O bO.~ ~ ~ ~ C O
~ a~
O O C ~
a a. o 3
v a) ~ ~ ~~ C o
_ o , ._, ~ 0._, a ~ ~
o u~~, c u~ ~6 ~ ~ c,
E~ a o E~ o~ o ~ ~
, o
v~ E ~J
u~ ~o z E~ o
~, -3~
2014882
, ~,~,,
The data in Table VI demonstrates the
effectiveness of the collector composition of the
present invention in the flot~tion of difficult to flGat
Arizona copper oxide ore.
Example 7 - Flotation of Mixed Oxide/Sulfide Copper Ore
A series of 30-g samples of -10 mesh copper ore
from Canada were prepared. The make-up of the valuable
portion of the ore was approximately 50 percent
malachite [Cu2C03(0H)2] and 50 percent chalcopyrite
[CuFeS2]. Each sample was ground along with 15 grams of
deionized water in a rod mill (2.5 inch diameter with
0.5 inch rods) for 1000 revolutions. The resulting pulp
was transferred to a 300 ml flotation cell. The pH of
the slurry was adjusted to 9.0 by the addition of sodium
carbonate. The hydroxy-containing compound, collector
and frother were added as described in the previous
examples.
The float cell was operated and samples were
prepared and analyzed as described in Example 6. The
results obtained are given in Table VII following.
3o
~,, 20148~2
ao a~ ~ ~ ~ ~D
~1
o o o o o C
O ~ -- ~ a
a) ~ ~ ~ o
o o o o o o
O
,1 ' '
o o o o
~ a~ In CT~ ao
c~ ~c) e-- I ~ I ~ u~
o , ~ , o o
o o o o
~V
O ~D ~ O
~ ~
~t C o o o o
,
t ~ ~ CO
--C~ ~ t~
o ~ ~
E~ O O C
o
I o o o o o o ~
c
~ ,
a), C O O O O O O O O O
~ o o o o o o o o o ~
~ o o o o o o o o o ~,
_ s~
~ ~ o O c ~) o ~ c~ ~o (~) o a~ c~ E
S c ~ d C~ C C
-- l E F ' E C-E~~ o 1:~
cl ~
5~
201~882
s/
~a.
The data in Table VII above generally
demonstrate the effectiveness of this invention in the
flot~tion of mixed copp~ ~xi~e/sulfide ores.
Example 8 - Flotation of Corundum
A series of 30-g samples of a -10 mesh mixture
of corundum (A1203) and silica (SiO2) were prepared.
Each sample was ground and transferred to a 300 ml
flotation cell as described in Example 7 with the
exception that the sample was ground 2000 revolutions.
The pH of the slurry was left at the natural pH of 7.4.
Collector, hydroxy compound and frother were added and
the float cell operated as described in Example 7.
Samples were obtained as described in Example 7 and were
dried, weighed, pulverized and the aluminum content
determined by X-ray fluorescence. The results obtained
are shown in Table VIII following.
3o
--- 201488~
-52- 64693-4628
t-- ~ Cl~ 3 t~J
O O
O '. ~ ~
O O O O O
o
E~ 3 c~ 3 ~o O
~:
O O O O O
V
C O
I ~o I In I ~
O O O
C
c~ ~ ~ J
n) ~O ~ î t-- î 3
a~ o I o I o
0: . I . I
E o o o
c
._,
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C ~ vl . ~ ¦ N
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_1
eS ~ ~ ~O
I C) ~r) I ~ I~D
O ~) I ~ 1 3 C
O O O
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~ g O g g g g C
,~ O O O O S
C O
u ~ _1 _l O ~ ~ O c ~ E
~ ~ o ' a O E
a c
~3 9 ~) 3 ~
~?9i~ '
~q
r ~ ~3 201~X82
The data shown in Table VIII above demonstrates
the effectiveness of the present inventiQn in the
separation G aLuminum from silica by flotation.
Example 9 - Flotation of Various Oxide Ores
The general procedure described in Example 1
were followed with the exception that various oxide ores
were used in place of the copper ore of Example 1. The
results obtained are shown in Table IX following.
3o
2014882
4~
TABLE IX
Recoveries of lifferent Minerals as a Function of
pH and Collector ~omposition Usin~ Propylene Glycol and
Oleic Aci~ at a Dosa~e of .012 k~/k~ Each
MINERAL pH 10.00
Pyrite, FeS2 1.000
Silica, SiO2 0.086
10Bauxite, Al(OH)3 0.913
Cassiterite, SnO2 1.000
Hematite, Fe203 1.000
Corundum, Al203 0.798
Calcite, CaCo3 1.000
Rutile, TiO2 1.0
Chromite, FeCr204 1.000
Dolomite, CaMg(CO3)2 1-000
Apatite, 1.000
20Ca5 ( C 1 lF ) [P04 ]3
Galena, PbS 1.000
Chalcopyrite, CuFeS2 1.000
Chalcocite, Cu2S 1.000
25Sphalerite, ZnS 1.000
Sylvite~ 0.703
Pentlandite, Ni(FeS)~ 1.000
Nickel Oxide (NiO) 0.911
~ Process carried out in saturated KCl solution
at pH 12.1.
Sample includes some pyrrhotite.
5~
2014882
This example demonstrates the efficacy of the
present invention in floating a broad range of oxide and
sulfide minerals. Also demonstrated is the ability io
distinguish these various minerals from silica, the
major gangue constituent found with these minerals in
natural ores.
Example lO
This example used the general Hallimond tube
procedure outlined in Example l except that instead of
using only pure mineral specimens in each run, a
specific test consisted of running a pre-mixed sample of
10 percent malachite (or 10 percent chrysocolla) along
with 90 percent silica. Copper assays were performed on
flotation concentrate and flotation tailings using the
acid dissolution procedure and D.C. plasma spectrometry
as discussed in Example 3. The results are shown in
Table Xa for malachite and Table Xb for chrysocolla.
All runs were determined at a pH of 10.0 with the
collector dosages as indicated.
3o
201~882
-56- 64693-4628
Table Xa
Malachite/Slllca Mlxture Separatlon
Do~age Cu
Collector (R~/k~) Recovery Cu Grade
Ethanol amine~ 0.024 0.078 0.799
Ethanol amlne 0.012 0.990 0.309
Oleic acid 0.012
Diethanol amlne~ 0.024 0.050~0.900
10Dlethanol amine 0.012 0.892 0.404
Oleic acid 0.012
~ Not an embodlment of the lnvention
3o
201488~
-57- 64693-4628
Table Xb
Chrysocolla/Slllca Mlxture Separatlon
Do~age Cu
Collector (Ka/ka) Recovery Cu Grade
Oleic acld~ 0.024 0.672 0.187
Oleic acld~ 0.012 o.389 0.324
~ Not an embodiment of the invention
3o
2Q148~2
.. ~
~,~
It is apparent from Tables Xa and Xb that a
number of hydroxy-containing compounds are effective in
decreasing the amount of silica gangue floated and
generally resulting in increased recovery and grade.
Example 11
A series of samples containing 30 g of a -10
mesh (U.S.) mixture of 10 percent rutile (TiO2) and 90
percent silica (sio2) were prepared. The remainder of
the procedure was exactly the same as that used in
Example 6.
- 201 1882
-59- 64693-4628
n o ~D
Vl ~
,_, o o o o o
o
J o
C)U~ O
a)o ~ o a~
Cl:
~ O O O O O
C~
C -- ~ O `O
,, ~ O O O O O
.) C
~, .~
~: I a)o o o o o
O O O O O
X ._1
-~ C
H n~ V I . -
C
~_ ~ . J J a~
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~ O O O O O
.~ ~
C
U ~
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~ o o o o o o o
C C O
c~ o o ~u O ~ ~ ~ E
O ~ ~ ~ ba a) r. r o au D
o ~" ~0 0 e
a a C
c~
cr;I ~ N tr) ~ U~ Z
2û14882
The data in Table XI above demonstrates the
effect of the present invention in increasing titanium
grade and recovery.
Examp~e 12 - Separation of Apatite and Silica
A series of 30-g samples of a -10 mesh (U.S.)
mixture of 10 percent apatite (Cas(Cl1F)[P04]3) and 90
percent silica (SiO2) were prepared. The remainder of
the procedure was exactly the same as that used in
Example 6. The natural ore slurry pH is 7.1.
~d
2014882
-
-61- 64693-4628
~
~1
_~ o o o
vO
E~ ~ u
C~
o~
oc
o o o
c u~ ~ ~
s.. I
V ~1 o O
~, ~ o o o
a~ c
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O I ¦ O~ O
,~ E oo o
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V
x ~ ~ v C.~l o o
o o
C
~ ~ c
¢ c ~ ~ ~ O
O
O ~ cr~ ~CJ~ ~
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v o o o ~
._, ~
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n ~ VD O OO O O O O c~
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~ O
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C) , ~ O ~ ,_ ~ ~ O ~ O
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201~882
s I ~)~r) ~ N
Vl. . . .
~ O O O O
J~
O
~ c~ ~ ~ a~ o
s a) cr~
O O O O
~ a) Vl j
~D O
a) O ._,
~ ~ In
~ X
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o
X cq O ~ V i
-1 ~ ~ o
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a.) I ~D00
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a ~- ~
a~
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~o
2~14882
- ~A
The data presented above demonstrates that the
use of hydroxy-containing compounds of this invention
witl1 oleic acid (which is a recognized collector for the
flotation of apatite) gives better grade and faster
flotation kinetics than the oleic acid alone. The
recoveries of apatite with all collectors was quite high
although slight improvements were observed in all cases
using the hydroxy-containing compounds of this
invention. Likewise, grade was improved in each case
with substantial improvement being shown in Run 1.
Example 13 - Flotation of Chalcopyrite Copper Ore
In this example, the effect of different
alkanol amines on the flotation of copper ore in
laboratory flotation cells was examined. Samples of
copper ore from Western Canada containing 500 g per
sample were prepared. The ore was relatively high grade
and also contained significant amounts of silica gangue.
A 500-g portion of the ore was ground with 257 g
deionized water in a rod mill having 2.5 cm rods at
about 60 revolutions per minute (RPM) for about 7
minutes. This produced a size distribution of 25
percent less than 100 me~h. Except as indicated in
Table I, the alkanol amine was added to the mill prior
to the grinding step. Lime was also added to the mill
to produce the de~ired pH for the subsequent flotation.
The pulp was transferred to a 1500-ml Agitair
Flotation cell outfitted with an automatic paddle
removal system. The ceii was agitated at 1150 RPM. The
pH of the slurry adjusted to 8.5 by the addition of
2014882
G
additional lime, if necessary. The collector, potassium
amyl xanthate unles~ specified otherwise in Table XIII,
was added to the slurry at a dosage of 8 g per ion an~
the slurry allowed to condition for one minute. A poly-
glycol ether frother, in the amount of 18 g per ton of
dry ore, was then added and the slurry allowed to
condition for one additional minute.
The flotation cell was agitated at 1150 RPM and
air introduced at a rate of 4.5 liters per minute.
Samples of the froth concentrate were collected for a
period of eight minutes after the air was first intro-
duced into the cell. These samples of the tailings and
concentrate were dried overnight in an oven, weighed,
and pulverized for analysis. After being pulverized,
they were dissolved with the use of acid and the copper
content determined using a DC Plasma spectrometer. The
assay data was used to determine fractional recoveries
and grades using standard mass balance formulas. The
recoveries represent the fractional amount of the
specified mineral present that was recovered.
Selectivity was determined by dividing the copper
recovery by the silica gangue recovery.
The data obtained is shown in Table XIII
following.
201~882
r,p~
TABLE XIII
Do~age Copper Silica
Alkanol (kg/met- Recov_ Recov- Selec-
Run Amine ric ton) er~ ery tivity
1~ None -- 0.654 0.135 4.8
2 Ethanol 0.020 0.663 0.114 5.8
amine
3 Diethanol 0.020 0.677 0.087 7.8
amine
4Triethanol 0.020 0.669 0.096 7.0
amine
5 Propanol 0.020 0. 673 0.118 5.7
amine
6Dipropanol 0.020 0. 683 0.093 7.3
amine
7 Isopro- 0.020 0.668 0.107 6.2
panol
amine
8 Butanol 0. 020 0.682 0.127 5. 4
1 0 amine
9 Diethanol 0. 040 0.648 0.079 8.2
amine
10~ Diethanol 0. 080 0.617 0.074 8.4
amine
11~Diethanol 0. 020 0.668 0.093 7.2
amine
12~Diethanol 0.040 0.627 0.089 7.2
amine
13~Diethanol 0. 020 0. 597 0.105 5.7
amine
14~Diethanol 0.040 0.544 0.095 5. 8
amine
2~14882
TABLE XIII, Continued
Dosage Copper Silica
Alkanol (kg/met- Recov- Recov- Selec-
Run Amine ric ton) er~ ery tivity
15~ None -- 0. 660 0.137 4.8
16~ None -- 0.582 0.128 4.5
17~Diethanol 0.020 0.658 0.100 6.6
amine
18~Diethanol 0.040 0.644 0.088 7.3
amine
19 Isopro- 0.040 0.649 0.095 6.8
panol
amine
20~Diethanol 0.020 0.658 0.117 5.6
amine
Not an embodiment of the invention.
N-ethyl isopropyl thionocarbamate used as collector.
~ Sec-butyl dithiophosphate used as collector.
In this run, the amine was added to the flotation cell
rather than grinding mill.
In this run, the amine and collector were added to the
flotation cell concurrently.
201~882
.
~! ~
The data in Table XIII demonstrates that the
practice of this invention is effective in decre~sing
the recovery of silica gangue and thus increasing the
selectivity of the flotation process. The data also
demonstrates t~at the practice of this invention can
result in lower recovery of the desired copper mineral
values. A comparison of Runs 3, 17 and 20 shows that
addition of the amine in the grinding stage rather than
in the flotation cell or concurrently with collector
results in the highest recovery of high grade copper.
Example 14 - Flotation of Mixed Copper Ore
A series of 30-g samples of mixed copper
sulfide ore from Nevada were prepared. The make-up of
the valuable components of the ore was about 0.25 weight
percent copper, about 0.004 weight percent molybdenum
and about 4 g/metric ton gold. Each sample of ore was
ground dry for about 20 seconds in a swing mill to about
12 percent greater than 100 mesh. The resulting ore
was transferred to a 300 ml flotation cell and diluted
with water.
The pH of the slurry was adjusted to 8.5 with
lime. The alkanol amine as specified in Table XIV was
added and the slurry allowed to condition for one
minute. Next, a first portion of the collector, sodium
isopropyl xanthate, (0.050 kg/metric ton of ore) was
added with an additional minute of conditioning. Next,
~7
~ 2014882
.
the frother, a polyglycol ether, was added in an amount
equivalent to 0.020 g per ton of dry ore and the slurry
allowed to condition an additional ~llinute.
The float cell was agitated at 1800 RPM and air
introduced at a rate of 2.7 liters per minute. Samples
o~ the froth concentrate were collected by standard hand
paddling at 2.0 minutes after the start of the
introduction of air into the cell. Next, a second dose
of collector (0.025 kg/metric ton of ore) was added with
one minute of conditioning an~ a six minute concentrate
collected. Samples of the concentrate and the tailings
were combined and then dried and analyzed as described
in the previous examples. The results obtained are
presented in Table XIV following. In each case, the
copper, gold, molybdenum and silica recoveries represent
the total amount recovered at the 2 and 6 minute
intervals.
TABLE XIV
Dosage Copper Gold denum Silica
Alkanol (kg/met- Recov- Recov- Recov- ecov-
Run Amine ric ton) ~ ery ery
1Diethanol amine 0.100 0.658 0.552 0.529 0.197
2Diethanol amine 0. 050 0.671 0.583 0.541 0.217
3Diethanol amine 0.200 0.614 0.529 0.498 0.183
4Monoethanol amine 0.100 0. 647 0.541 0.511 0.209
5Triethanol amine 0.100 0.653 0.557 0.518 0.213
6I~opropanol amine 0.100 0.651 0.549 0.523 0.217
7~ None -- 0. 624 0.533 0.489 0.250
Not an embodiment of the invention.
2014882
, ~
~,
The data shown above demonstrates the
effectiveness of the process of the present invention in
increasing Ihe grade of recovered minerai va7ues.
Example 15 - Flotation of Mixed Sulfide/Oxide Copper Ore
The general procedure outlined in Example 13
was followed using a southern Africa mixed sulfide/oxide
copper ore. The sulfide copper ore was floated by the
practice of this invention and the remaining oxide ore
recovered in a subsequent step such as leaching or oxide
flotation. The sulfide minerals contained in this ore
was quite small, less than about 0.22 weight percent of
the total ore.
One modification to the procedure outlined in
Example 13 was that the ore was ground for 700
revolutions to produce a size distribution of 13 percent
greater than 100 mesh. The collector used was potassium
amyl xanthate at a concentration of 0.025 kg/metric ton
of ore. In each case, the alkanol amine used was
diethanol amine in the amounts specified. The results
obtained are shown in Table XV following.
-IA 201488~
TABLE XV
Dosage CopperLead Zinc Silica
(kg/met- Recov- Recov- Recov- Recov-
Run ric ton) ery ery ery ery
1~ None 0.704 0.835 0.491 0.317
2 0.025 0.714 0.831 0.486 0.273
3 0.050 0.693 0.824 0.480 0.246
4 0.100 0.650 0.791 0.452 0.209
0.200 0.589 0.746 0.396 0.152
Not an embodiment of the invention.
The data above again show that the practice of
the present invention results in decreasing recoveries
of silica gangue. With this particular ore, the
recovery of the desired mineral values of lead and zinc
20 also declined even at the lowest dosage of the alkanol
amine.
Example 16 - Effect of Order and Manner of Addition of
Collector and Hydroxy-Containing Compound
The procedure outlined in Example 6 was
followed with the exception that the apatite used was
from a dif~erent source and contained about 30 percent
30 apatite and about 70 percent silica. The hydroxy-
-containing compound used in each case was diethanol
amine and the anionic collector oleic acid. In each
runj the manner in which the diethanol amine a..d oleic
acid were added to the flotation system varied. In Run
1, diethanol amine was added to the cell and allowed to
7/
~o2 2014882
condition for one minute. This was followed by the
addition of the oleic acid followed by an additional
minute of conditioning. In Run 2, the order of addition
is reversed. In Run 3, diethanol amine and oleic acid
were each added to the cell at the same time and in
approximately the same physical location and a lowed to
condition for one minute. In Run 4, diethanol amine and
oleic acid were mixed in a separate container and a salt
was formed as indicated by the evolution of heat. This
was added to the flotation cell and then conditioned for
one minute. In Run 5, a condensate of exccss fatty
acids and diethanol amine available commercially as M-
210 from The Dow Chemical Company was used in place of
unreacted oleic acid and diethanol amine. In Runs 6
and 7, oleic acid was used alone. The results obtained
are shown in Table XVI following.
2014882
J ~ O t~
~ ~ N ~ D 0 0
C7 1 ' ~ '~ ~ O O ~
~ O O O O O O O
q) O
cO a~ C-- L~ J N
O N N ~ O O
a) a~ o~ ao ~ O ~ a~
O O O O O O O
r
t~ ~ ~ ~ J ~ r-- ~
0 U~ CJ~ N ~0
~ l . o o o
a~3 0 0 0 O O O
a) O._,
O
3 a~ O N ~ J ~ ~~n
N J N
x , a~ o o o o o o
--I v2 N ~
S ~ O O O O O O O
t~ c
~ S J
H , ~ N N ~ N~O ~ ~
~ ~ ~ O O O O O O O
m s~ o
'~ N o l o ~ o o D a~
E~ a) I ~ O~ ~ o
~ o ~; . . . . . . . a
-~ o o o o o o o
Cd ._
Q I ,_
O O O O O O O O S.
~70 O O ~ . - N ON NO ,O J~
~.~ OOOO OOO OOOo
r C ~ C ~
~ o O a a O a E ~ C~ 3 o
r~
N ~ J In ~
~,4 2014882
Runs 1-3, embodiments of this invention clearly
demonstrate its effectiveness. Run 4 shows that when
the components of the invention are pre-mixed, the
recovery of phosphorus obtained is substantially less
than when oleic acid is used alone. Run 5 shows that a
fatty acid/diethanol amine condensate is ineffective in
this process.
