CA2014882C - Depression of the flotation of silica or siliceous gangue in mineral flotation - Google Patents

Abstract

A flotation process is disclosed wherein the selectivity to the valuable mineral is improved by the depression of silica or siliceous gangue by the use of a hydroxy-containing compound such as an alkanol amine.
The process is useful in the flotation of oxide minerals as well as in sulfide flotation.

Description

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 ~

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 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 -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 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 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.

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);

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), 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 _ 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 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 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+

-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 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 _ -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:

~ P-Z--M+

~ 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 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:

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

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 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.

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 , 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 .~

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 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 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 -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

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 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 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 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 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 . 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

~, ~

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 201~882 ~ 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 ~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~
, ~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 -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 . ~ ~ 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~

, ~,~,, 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~

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 ._, E o I J
C ~ vl . ~ ¦ N
C
_1 eS ~ ~ ~O
I C) ~r) I ~ I~D
O ~) I ~ 1 3 C
O O O
C
~ 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 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~

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 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~

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.

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~
~1 ~ o t- ~ c ~ O O O O O
.~ ~
C
U ~
~' E " O o o ~C
~ 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.

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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 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 .
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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.

, ~
~, 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 ~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.

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~,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.

Claims (6)

1. 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 using an anionic collector 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 lower alkanol amine selected from ethanol amine, propanol amine, butanol amine, diethanol amine, dipropanol amine, tripropanol amine, triethanol amine and mixtures thereof and the collector and the lower alkanol amine are not premixed prior to being added to the process.

2. The process of Claim 1 wherein the anionic collector is derived from an acid selected from carboxylic, sulfonic, sulfuric, phosphoric and phosphonic acids.

3. The process of Claim 2, wherein the anionic collector comprises alkyl sulfonic acids, alkylaryl sulfonic acids, their salts and mixtures thereof.

4. The process of Claim 3, wherein the anionic collector is selected from alkylated benzene sulfonic acid, alkylated sulfonic acid, alkylated diphenyl oxide monosulfonic acids, their salts and mixtures thereof.

5. The process of Claim 3, wherein the anionic collector is selected from linolenic acid, oleic acid, lauric acid, linoleic acid, octanoic acid, capric acid, myristic acid, palmitic acid, stearic acid, arachidic acid, behenic acid, 2-naphthalenic sulfonic acid, sodium lauryl sulfate, sodium stearate, dodecane sodium sulfonic acid, dodecyl sodium sulfate, dodecyl phosphate, chloride derivative of dodecyl phosphonic acid, 2-naphthoic acid, pimelic acid, 11-aminododecanoic acid, dedecyl benzyl sulfonic acid, hexadecyl sulfonic acid, their salts and mixtures thereof.

6. The process of Claim 1 including the use of a thiol collector selected from thiocarbonates, thionocarbamates, thiocarbanilides, thiophosphates, thiophosphinates, mercaptans, xanthogen formates, xanthic esters and mixtures thereof.