WO1995003891A1

WO1995003891A1 - Aryl ether monosulfonate collectors useful in the flotation of minerals

Info

Publication number: WO1995003891A1
Application number: PCT/US1994/008532
Authority: WO
Inventors: Richard R. Klimpel; Kevin A. Frazier
Original assignee: Dow Chemical Co
Current assignee: Dow Chemical Co
Priority date: 1993-07-29
Filing date: 1994-07-26
Publication date: 1995-02-09
Anticipated expiration: 1996-01-29
Also published as: FI943509A0; AU7476194A; FI943509A7; FI943509L; TR28785A; ZA945513B; PE6695A1

Abstract

A flotation process is disclosed using as a collector an alkylaryl alkyl ether sulfonic acid, a monoalkylated phenol sulfonic acid or an alkylene oxide derivative of an alkylated phenol sulfonic acid. Preferred collectors include C16-24 monoalkylated phenol sulfonic acids and salts thereof.

Description

ARYL ETHER MONOSULFONATE COLLECTORS

USEFUL IN THE FLOTATION OF MINERALS

This invention is related to the use of chemical collectors in the recovery of minerals by froth flotation.

A wide variety of chemical reagents are recognized by those skilled in the art as having utility in froth flotation. However, it is 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 comprised of oxygen-containing species such as carbonates, hydroxides, sulfates and silicates.

While a large proportion of the minerals existing today are contained in oxide ores, most successful froth flotation systems are directed to sulfide ores. Most known recovery methods have not been economically feasible for the recovery of oxide ores and,

consequently, a large proportion of oxide ores are simply not processed. Thus, the need for improved oxide mineral flotation is generally acknowledged by those skilled in the art of froth flotation

This invention is a process for the recovery of minerals by froth flotation characterized by the use of a collector comprising a sulfonic component, and

optionally other collector components, the sulfonic component selected from alkylaryl alkyl ether sulfonic acids or salts thereof; mono C_16-24 alkylated phenol sulfonic acids or salts thereof; and alkylene oxide derivatives of C_16-24 alkylated phenol sulfonic acids or salts thereof.

The recovered minerals may be the minerals that are desired or may be undesired contaminants.

Additionally, the froth flotation process of this invention may utilize other collectors, frothers and other flotation reagents known in the art.

Particular advantages of the collectors of the present invention include their effectiveness at

relatively low dosages and the relative simplicity of their manufacture.

Non-limiting examples of oxide ores which may be floated using the practice of this invention

preferably include iron oxides, nickel oxides, copper oxides, phosphorus oxides, aluminum oxides and titanium oxides. Other types of oxygen-containing minerals which may be floated using the practice of this invention include carbonates such as calcite or dolomite and hydroxides such as bauxite.

Non-limiting examples of specific oxide ores which may be collected by froth flotation using the process of this invention include those containing 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, topaz and samarskite.

The process of this invention is also useful in the flotation of sulfide ores. Non-limiting examples of sulfide ores which may be floated by the process of this invention include those containing chalcopyrite,

chalcocite, galena, pyrite, sphalerite, molybdenite 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, may also be recovered by the practice of this invention. For example, such metals are sometimes found associated with oxide and/or sulfide ores. Platinum, for example, may be found associated with troilite. By the practice of the present

invention, such metals may be recovered in good yield.

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 small amounts of noble metals as discussed above. Minerals may be recovered from these mixed 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 and collector composition of the present invention to recover primarily oxide minerals and any noble metals that may be present.

Alternatively, both the sulfur-containing and oxygen-containing minerals may be recovered

simultaneously by the practice of this invention.

The selectivity demonstrated by the collectors of this invention permits the separation of small amounts of undesired minerals from the desired minerals. For example, the presence of apatite is frequently a problem in the flotation of iron as is the presence of topaz or tourmaline in the flotation of cassiterite. Thus, the collectors of the present invention are, in some cases, useful in reverse flotation where the undesired mineral is floated such as floating topaz or tourmaline away from cassiterite or apatite from iron.

In addition to the flotation of ores found in nature, the flotation process and collector composition of this invention are useful in the flotation of

minerals from other sources. One such example is the waste materials from various processes such as heavy media separation, magnetic separation, metal working and petroleum processing. These waste materials often contain minerals that may be recovered using the

flotation process of the present invention. The compounds useful as collectors in the practice of this invention include alkylaryl alkyl ether sulfonic acids and salts thereof; mono C_12-24 alkylated phenol sulfonic acids and salts thereof; and alkylene oxide derivatives of alkylated phenol sulfonic acids and salts thereof. These compounds preferably correspond to the following formula:

wherein Ar is benzene, napthalene, anthracene and compounds corresponding to the formula:

wherein X represents a covalent bond (i.e., biphenyl); ╌(CO)╌; or R³ wherein R³ is a linear or branched alkyl divalent moiety having one to three carbon atoms; R¹ is a C_1-30 linear or branched alkyl group; R² is hydrogen, a C_1-30 linear, branched or cyclic alkyl group or

((-CH₂CHR⁴O)_aH wherein a is 1-6, preferably 1 or 2 and R⁴ is hydrogen, methyl or ethyl; and M is independently hydrogen, an alkali metal, alkaline earth metal, or ammonium or substituted ammonium, with the proviso that the total carbon content of R¹ and R² is from 16 to 34 carbon atoms, preferably 16 to 24 carbon atoms. The M⁺ ammonium ion radicals are of the formula (R')₃HN⁺ wherein each R' is independently hydrogen, a C₁-C₄ alkyl or a C₁-C₄ hydroxyalkyl radical. Illustrative C₁-C₄ alkyl and hydroxyalkyl radicals include methyl, ethyl, propyl, isopropyl, butyl, hydroxymethyl and

hydroxyethyl. Typical ammonium ion radicals include ammonium (N⁺H₄), methylammonium (CH₃N⁺H₃), ethylammonium (C₂H₅N⁺H₃), dimethylammonium ((CH₃)₂N⁺H₂), methylethylammonium (CH₃N⁺H₂C₂H₅), trimethylammonium ((CH₃)₃N⁺H), dimethylbutylammonium ((CH₃)₂N⁺HC₄H₉),

hydroxyethylammonium (HOCH₂CH₂N⁺H₃) and

methylhydroxyethylammonium (CH₃N⁺H₂CH₂CH₂OH).

It is preferred that Ar is a benzene ring. In one preferred embodiment, R¹ is a C_16-24 linear or branched alkyl group and R² is hydrogen or a C_1-4 linear or branched alkyl group. R¹ is preferably linear and attached to Ar through a carbon atom other than a terminal atom. M is preferably hydrogen, sodium, calcium, potassium or ammonium. It should be noted that if R¹ is lower alkyl, i.e., C_1-4, then R² must be C₁₂ or greater so that the total carbon content is from C_16-24.

In a second preferred embodiment, R¹ is a

C_16-24 linear or branched alkyl group and R² is

((-CH₂CHR⁴O)_aH wherein a is is 1-6, preferably 1 or 2 and

R⁴ is hydrogen, methyl or ethyl. M is preferably hydrogen, sodium, calcium, potassium or ammonium.

As shown in the formula above, the collectors useful in the practice of this invention have a single sulfonation. Those skilled in the art will recognize that when Ar is an aromatic group having more than one ring, it is possible to have additional sulfonations, depending on the manner in which the molecule is

sulfonated. While some proportion of multiple

sulfonation is acceptable, a substantial portion of the molecules of the collector of the present invention must be monosulfonated. By substantial portion, it is meant that at least 20 percent, more preferably at least

35 percent and most preferably at least 50 percent of the collector is monosulfonated. It should be noted that the collectors of this invention may be in acid or salt form and, in general, any reference to sulfonic acid includes the sulfonate and any reference to the sulfonate includes the sulfonic acid.

Examples of alkylaryl alkyl ether sulfonic acids useful as collectors include octadecyl anisole sulfonic acid, C_20-24 alkylated anisole sulfonic acid, C_24-28 alkylated anisole sulfonic acid, dodecyl anisole sulfonic acid, hexadecylphenyl butyl ether sulfonic acid, octadecylphenyl butyl ether sulfonic acid, C_20-24 alkylphenyl butyl ether sulfonic acid, octadecyloxy cumene sulfonic acid, octadecyloxy toluene sulfonic acid, C_18-24 alkylatedphenyl isopropyl ether sulfonic acid, sulfonic acid of C_20-24 alkylphenol, sulfonic acid of octadecylphenol, sulfonic acid of hexadecylphenol, and sulfonic acid of dodecylphenol. It is preferred that the collector of the present invention include alkylated anisole sulfonic acid, more preferably

octadecyl anisole sulfonic acid.

When the collector is an alkylated phenol sulfonic acid or salt thereof, it is preferably sulfonic acid of octadecylphenol or sulfonic acid of

hexadecylphenol.

When the collector is an alkylene oxide derivative of an alkylated phenol sulfonic acid or salt thereof, it is preferred that it corresponds to the formula:

wherein R⁵ _n is a C₂₀ to C₂₄ alkyl group, n is 1 or 2 and R⁶ is ╌(CH₂CH₂O)_mH wherein m is 1 or 2; or wherein R⁵ is a C₁₂ to C₁₆ alkyl group and R⁶ is ╌(CH₂CHR⁷O)_mH wherein R⁷ is methyl or ethyl, preferably methyl and m is 1 or 2

The alkylaryl alkyl ether sulfonic acids or salts thereof of this invention may be made by

techniques known to those skilled in the art. For example, materials such as anisole or phenetole may be alkylated on the aromatic ring via Friedel-Crafts chemistry using alkylating agents such as alpha-olefins, alkylhalides or alkyl alcohols. Typical catalysts include Lewis acids, mineral acids and the acid form of sulfonated polystyrene beads. Alternatively, the ether may be prepared by 0-alkylating an alkylated phenol with an alkyl halide under basic conditions, either prior or post sulfonation. The alkylaryl alkyl ether may be sulfonated either prior to or following alkylating the aryl ring, but it is more practical to perform the sulfonation step last. Sulfonation can be accomplished using established methods such as through the use of sulfur trioxide in methylene chloride, sulfur trioxide in air, or chlorosulfonic acid in methylene chloride. The alkylated phenol sulfonic acid or salt thereof may be prepared by sulfonation of the corresponding alkylated phenol. Sulfonation can be accomplished using established methods such as through the use of sulfur trioxide in methylene chloride, sulfur trioxide in air, or chlorosulfonic acid in methylene chloride.

The alkylene oxide derivative of an alkylated phenol sulfonic acid or salt thereof may be prepared by known methods. For example, the corresponding alkyl phenol may be capped with the appropriate alkylene oxide equivalents prior to sulfonation using known methods.

The collector of this invention may be used as the only collector or it may be used in conjunction with other collectors. One example where the collector of this invention is not used alone as a collector is when it is used in combination with a carboxylic component. The carboxylic component is a C_1-24 carboxylic acid or salt thereof. Examples of useful materials include acetic acid, citric acid, tartaric acid, maleic acid, oxalic acid, ethylenediamine dicarboxylic acid,

ethyleneamine tetracarboxylic acid and fatty acids.

Fatty acids or their salts are particularly preferred. Illustrative examples of such acids include oleic acid, linoleic acid, linolenic acid, myristic acid, palmitic acid, strearic acid, palmitoleic acid, caprylic acid, capric acid, lauric acid and mixtures thereof. One example of a mixture of fatty acids is tall oil.

Preferred fatty acids include oleic acid, linoleic acid, linolenic acid and mixtures thereof. The fatty acids may be used in the acid form or may be used in salt form. As used herein, the terms "acid" and"carboxylate" include both the acid and salt form. When used, the amount of carboxylic component is preferably at least 1 weight percent, more preferably at least 2 weight percent and most preferably at least 5 weight percent, based on the combined weight of the sulfonic acid or salt and the carboxylic component. The maximum amount of carboxylic component used is

preferably no greater than 50 weight percent, more preferably no greater than 40 weight percent, and most preferably no greater than 30 weight percent. As will be recognized by one skilled in the art, the optimum amount of carboxylic component used depends on the degree of hardness of the water used in flotation, the minerals to be recovered and other variables in the flotation process.

The carboxylic component may be added to the flotation system prior to the addition of the collector of the present invention or they may be added

simultaneously. In one embodiment, the sulfonic

component and the carboxylic component are formulated and then added to the flotation system. The collector composition may be formulated in a water based mixture or a hydrocarbon based mixture, depending on the

particular application. When a water formulation is used, the sulfonic component and/or the carboxylic component are in the salt form. When a hydrocarbon based formulation is used, one or both of the components is in the acid form. Typical hydrocarbon formulations include any saturated hydrocarbon, kerosene, fuel oil, alcohol, alkylene oxide compound, or organic solvents such as dodecene, dimethylsulfoxide, limonene and dicyclopentadiene. In preferred embodiments, both the sulfonic and carboxylic components are in either the salt form or the acid form. Mixed formulations where one is a salt and the other an acid are possible, but are generally not preferred. The acid form, or hydrocarbon based

formulations, are generally preferred in those

situations where pH regulators are used to raise the pH above 7. In those instances where the flotation was conducted at a natural pH, it is typically preferred to use the salt form or water based formulations.

The collector can be used in any concentration which gives the desired selectivity and recovery of the desired mineral values. In particular, the concentration used is dependent upon the particular mineral to be recovered, the grade of the ore to be subjected to the froth flotation process and the desired quality of the mineral to be recovered. A particular advantage of the present invention is the effectiveness of the collector at low dosage levels.

Additional factors to be considered in determining dosage levels include the amount of surface area of the ore to be treated. As will be recognized by one skilled in the art, the smaller the particle size, the greater the surface area of the ore and the greater the amount of collector reagents needed to obtain adequate recoveries and grades. Typically, oxide mineral ores must be ground finer than sulfide ores and thus require very high collector dosages or the removal of the finest particles by desliming. Conventional processes for the flotation of oxide minerals typically require a desliming step to remove the fines present and thus permit the process to function with acceptable collector dosage levels. The collector of the present invention functions at acceptable dosage levels with or without desliming.

Preferably, the concentration of the collector is at least 0.001 kg/metric ton, more preferably at least 0.05 kg/metric ton. It is also preferred that the total concentration of the collector is no greater than 5.0 kg/metric ton and more preferred that it is no greater than 2.5 kg/metric ton. In general, to obtain optimum performance from the collector, it is most advantageous to begin at low dosage levels and increase the dosage level until the desired effect is achieved. While the increases in recovery and grade obtained by the practice of this invention increase with increasing dosage, it will be recognized by those skilled in the art that at some point the increase in recovery and grade obtained by higher dosage is offset by the

increased cost of the flotation chemicals. It will also be recognized by those skilled in the art that varying collector dosages are required depending on the type of ore and other conditions of flotation. Additionally, the collector dosage required has been found to be related to the amount of mineral to be collected. In those situations where a small amount of a mineral susceptible to flotation using the process of this invention is present, a very low collector dosage is needed due to the selectivity of the collector.

It has been found advantageous in the recovery of certain minerals to add the collector to the

flotation system in stages. By staged addition, it is meant that a part of the collector dose is added; froth concentrate is collected; an additional portion of the collector is added; and froth concentrate is again collected. The total amount of collector used is preferably not changed when it is added in stages. 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.

An additional advantage of staged addition is related to the ability of the collector of the present invention to differentially float different minerals at different dosage levels. At low dosage levels, one mineral particularly susceptible to flotation by the collector of this invention is floated while other minerals remain in the slurry. At an increased dosage, a different mineral is floated thus permitting the separation of different minerals contained in a given ore.

In addition to the collector of this invention, other conventional reagents or additives may be used in the flotation process. Examples of such additives include various depressants and dispersants well known to those skilled in the art. Additionally, the use of hydroxy-containing compounds such as alkanol amines is useful in improving the selectivity to the desired mineral values in systems containing silica or siliceous gangue. In addition, frothers are preferably used.

Frothers are well known in the art and reference is made thereto for the purposes of this invention. Examples of useful frothers include polyglycol ethers and lower molecular weight frothing alcohols. Additionally, the collectors of this invention may be used with a

hydrocarbon as an extender. Examples of hydrocarbons useful in this context include those hydrocarbons typically used in flotation such as fuel oil, kerosene and motor oil.

As discussed above, the collectors of this invention may also be used in conjunction with other collectors. It has been found that in the flotation of sulfide mineral containing ores, the use of the

collector of this invention with sulfide thiol

collectors such as xanthates, dithiol phosphates and trithiol carbonates is advantageous.

The collectors of this invention may also be used in conjunction with other conventional collectors in other ways. For example, the collectors of this invention may be used in a two-stage flotation in which the collector of this invention recovers primarily oxide minerals while a second stage flotation using

conventional collectors recovers primarily sulfide minerals or additional oxide minerals. When used in conjunction with conventional collectors, a two-stage flotation may be used wherein the first stage comprises the process of this invention and is done at the natural pH of the slurry. The second stage involves

conventional collectors and is conducted at an elevated pH. It should be noted that in some circumstances, it may be desirable to reverse the stages. Such a two-stage process has the advantages of using less

additives to adjust pH and also permits a more complete recovery of the desired minerals by conducting flotation under different conditions.

A particular advantage of the collector of the present invention is that additional additives are not required to adjust the pH of the flotation slurry. The flotation process utilizing the collector of the present invention operates effectively at typical natural ore pH's ranging from 5 or lower to 9. This is particularly important when considering the cost of reagents needed to adjust slurry pH from a natural pH of around 7.0 or lower to 9.0 or 10.0 or above which is typically

necessary using conventional carboxylic xanthic

collectors. As noted above, a collector composition comprising the collector of the present invention and a xanthate collector is effective at a lower pH than a xanthate collector used alone.

The ability of the collector of the present invention to function at relatively low pH means that it may also be used in those instances where it is desired to lower the slurry pH. The lower limit on the slurry pH at which the present invention is operable is that pH at which the surface charge on the mineral species is suitable for attachment by the collector.

Since the collector of the present invention functions at different pH levels, it is possible to take advantage of the tendency of different minerals to float at different pH levels. This makes it possible to do one flotation run at one pH to optimize flotation of a particular species. The pH can then be adjusted for a subsequent run to optimize flotation of a different species thus facilitating separation of various minerals found together. The following examples are provided to illustrate the invention and should not be interpreted as limiting it in any way. Unless stated otherwise, all parts and percentages are by weight. The following examples include work involving Hallimond tube flotation and flotation done in

laboratory scale flotation cells. It should be noted that Hallimond tube flotation is a simple way to screen collectors, but does not necessarily predict the success of collectors in actual flotation. Hallimond tube flotation does not involve the shear or agitation present in actual flotation and does not measure the effect of frothers. Thus, while a collector generally must be effective in a Hallimond tube flotation if it is to be effective in actual flotation, a collector

effective in Hallimond tube flotation is not necessarily 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 required in an actual flotation cell.

Example 1 - Preparation of Collectors

Octadecyl Anisole Sulfonic Acid (IV) Anisole (108 g) and aluminum chloride (0.8 g) were heated together at 70°C for about 1 hour.

0ctadec-1-ene (25.2 g) was added dropwise during the first 1.25 hours. After about 1 hour the temperature of the mixture was increased to 85°C. After a total of about 2 hours of heating about 1 g of additional

aluminum chloride was added. The mixture was then cooled to room temperature and 25 ml 5N sodium hydroxide was added. This solution was stirred for 1 hour and then diluted with 150 ml hexane. The aqueous layer was then removed and the solution was washed once again with 25 ml of 5N sodium hydroxide. Then the solution was washed once with 1N hydrochloric acid and then once with saturated sodium chloride solution. The organic layer was dried overnight and then distilled under reduced pressure. The fraction which distilled at 160°-192°C at 0.005 mm Hg was collected and used for sulfonation.

Two grams of octadecyl anisole was added slowly to 20 percent fuming sulfuric acid which was stirring in an ice bath. After the addition was complete, the mixture was capped and stored in a freezer at -20°C for about 1-1/2 hours. Anhydrous ether (4 ml) was added to the cold mixture and stirred vigorously for 4 to

5 minutes. Hexane (40 ml) was added and the mixture was allowed to settle into two layers after an additional

5 minutes of stirring. The upper layer was decanted and collected. The addition of hexane, stirring and

decantation were repeated four more times. The combined hexane fractions were evaporated under reduced pressure, producing a product which appeared as a dark brown oil which eventually solidified to a dark brown wax.

Similar procedures were followed to prepare alkylated anisole sulfonic acid with the alkyl group being a C_20-C₂₄ alkyl; a C_24-C₂₈ alkyl; a dodecyl group; and a hexadecyl group.

Octadecylphenyl Butyl Ether Sulfonic Acid

Phenol (25.5 g), octadec-1-ene (63 g) and dry, strong cationic ion exchange resin (protonated form) (2.5 g) were heated at about 100°C for 24 hours. The resin was removed by filtration and the product

(45.95 g) was distilled under vacuum. The octadecylphenol prepared as described above was mixed with 1-bromobutane and 50 percent sodium hydroxide solution were mixed in N,N-dimethylformamide. This mixture was heated to 50°C for 48 hours and then cooled and transferred to a separatory funnel where it was diluted with hexane and washed three times with water and 5N sodium hydroxide. The organic layer was evaporated under reduced pressure. The octadecylphenyl butyl ether so obtained was sulfonated as described above.

This technique was also used to prepare an alkylphenyl butyl ether sulfonic acid wherein the alkyl group was a C_20-C₂₄ alkyl.

Octadecyloxy Cumene

Isopropylphenol (13.6 g), sodium hydroxide and octadecyl bromide (30 g) were mixed in 10 ml in water and heated to 100°C. Sodium hydroxide (3.8 g) dissolved in 4 ml of water was added over a period of about

10 minutes. The mixture was then heated an additional 4 to 5 hours. About 1 additional gram of isopropylphenol and about 1/2 additional gram of sodium hydroxide were added and the heating was continued for an additional 24 hours. The mixture was cooled, diluted with ether and washed two times with 5N sodium hydroxide, one time with a salt solution and then dried over sodium sulfate. Solvent was removed under reduced pressure yielding about 33.5 g octadecyloxy cumene. Sulfonation was conducted using 30 percent fuming sulfuric acid.

Octadecyloxy toluene sulfonic acid was prepared in a manner similar to that used for the preparation of octadecylphenyl butyl ether sulfonic acid. In a similar manner a C_18-C₂₄ alkylphenyl isopropyl ether sulfonic acid, an isopropylphenyl octadecyl ether sulfonic acid, a nonylphenyl tridecyl ether sulfonic acid and

decylphenyl butyl ether sulfonic acid were prepared. n-Octadecylphenol Butyl Ether Sulfonic Acid

Phenyl butyl ether (7.5 g) and aluminum chloride (6.67 g) were dissolved in 100 m of hexane. Stearoyl chloride (15.15 g) dissolved in 25 ml of hexane was added dropwise to the mixture. During the course of the addition temperature of the mixture rose from room temperature to about 45°C. After about 2.5 hours, the mixture was poured slowly into a mixture of ice and water and a sticky tan solid formed after vigorous stirring. This solid dissolved upon the addition of about 75 ml of hexane and 200 ml of THF

(tetrahydrofuran). The organic layer was separated and dried and the solvent was stripped under reduced

pressure. The product stearoyl phenyl butyl ether was used without further purification.

This product was mixed with 27.4 g of

trifluoroacetic acid and 6.14 g triethylsilane in 100 ml of hexane. The mixture was heated to 50°C for about four days. All by-products and solvents were removed leaving a residue, octadecylphenyl butyl ether, which was used without further purification.

Sulfonation was conducted using fuming sulfuric acid as described above. C₂₀-C₂₄ Alkyl Phenol, Sulfonic Acid, Sodium

Salt

Phenol (47 g), dry, strong cationic ion exchange resin (protonated form) (2.5 g), and a C_20-24 olefin were heated at 105°C for several hours under a nitrogen stream. The reaction mixture was cooled and filtered. The resin was washed with anhydrous ether to remove all the reaction materials from the resin and the ether and excess phenol were removed at reduced

pressure. The product obtained, which was the alkylated phenol, was dissolved in CH₂Cl₂ (30 ml) under a nitrogen atmosphere. Chlorosulfonic acid (0.80 ml) was added via a syringe. This addition caused an evolution of heat which caused the solvent to reflux. The remaining solvent was removed under reduced pressure and 2.4 ml of 5N sodium hydroxide was added. Ether was added to the mixture and it was stirred vigorously. The ether was removed under reduced pressure and the remaining

material was dried under vacuum overnight.

In a similar manner the following collectors were prepared: octadecylphenol, sulfonic acid, sodium salt; hexadecylphenol, sulfonic acid, sodium salt;

dodecylphenol, sulfonic acid, sodium salt.

Octadecyl β-Hydroxyethoxybenzene Sulfonic Acid, Sodium Salt

Octadecylphenol ( 5 g, 14.4 mmol), ethylene carbonate (1,27 g, 14.4 mmol) and potassium fluoride (0.25 g) were mixed together under nitrogen and heated to 180°C for 4 to 5 hours. The mixture was cooled and hexane (150 ml) was added. The solution was then washed three times with water. The organic layer was dried and evaporated under reduced pressure.

Crude octadecyl β-hydroxyethoxybenzene (3 g, 7.68 mmol) was dissolved in CH₂Cl₂ (20 ml) and cooled to 0°C under a stream of nitrogen. Chlorosulfonic acid (1.02 ml, 15.36 mmol) was added slowly over a period of about 5 minutes. The mixture was stirred slowly and after 4 hours, solvent was removed under reduced

pressure. Water (20 ml) was added to the thick, light brown product, followed by 5 N sodium hydroxide solution (3.1 ml, 15.5 mmol). The solution was heated at reflux overnight. Water was removed under reduced pressure.

Octadecyl β-hydroxypropoxybenzene sulfonic acid, sodium salt was prepared in a similar manner.

Table I below identifies the collectors prepared as discussed above and used in the following examples.

Example 2 - Flotation of Various Oxide Minerals

About 1.1 g sample of the specified mineral sized to about -60 to +120 U.S. mesh was placed in a small bottle with 20 ml of deionized water. The mixture was shaken 30 seconds and then the water phase

containing some suspended fine solids or slimes was decanted. This desliming step was repeated several times. A 150-ml portion of deionized water was placed in a 250-ml glass beaker. Next, 2.0 ml of a 0.10 molar solution of potassium nitrate was added as a buffer electrolyte. The pH was adjusted to the specified level with the addition of 0.10 N HCl and/or 0.10 N NaOH.

Next, a 1.0-g portion of the deslimed mineral was added along with deionized water to bring the total volume to about 180 ml. The specified collector was added and allowed to condition with stirring for 15 minutes. The pH was monitored and adjusted as necessary using HCl and NaOH. All collectors indicated were converted to the Na⁺ salt form before addition. The collector dosage in all tests was 1.0 g per 1.0 kg of pure mineral. 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 inches 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 100°C. Each portion was weighed and the fractional recoveries of each mineral were reported in Table II below. 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 recovery of each mineral reported was that fractional portion of the original mineral placed in the Hallimond tube that was recovered. Thus, a recovery of 1.00 indicated that all of the material was recovered. The values given for the various mineral recoveries generally were correct to +0.05. All runs were

conducted at a pH of 8.0. The collectors used are octadecyl anisole sulfonic acid (IV in Table I), the sodium salt of octadecylphenol, sulfonic acid (VI in Table I), and the octadecylphenol capped with one ethylene oxide unit (XIII in Table I). The results are shown in Table II below.

Example 3 - Hallimond Tube Flotation of Apatite

Using the procedure set out in Example 2, the ability of Collectors I through XIV (from Table I) to float apatite in a Hallimond tube was measured. The results obtained are shown in Table III below.

Example 4 - Separation of Apatite and Silica

A series of 30 g samples of a -10 mesh (U.S.) mixture of 10 percent apatite (Ca₅(Cl,F)[PO₄]₃) and 90 percent silica (SiO₂) was prepared. 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) 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 6.7. After addition of the collector (in the sodium salt form) as shown in Table IV, the slurry was allowed to condition for 1 minute. Next, the frother, a polyglycol ether available commercially from The Dow

Chemical Co. as Dowfroth^® 420 brand frother, was added in an amount equivalent to 0.050 kg per ton of dry ore and the slurry was allowed to condition an additional minute.

The float cell was agitated at 1800 RPM and air was introduced at a rate of 2.7 liters per minute. The froth concentrate was collected by standard hand paddling for 4 minutes after the start of the

introduction of air into the cell. Samples of the concentrate and the tailings were dried and analyzed as described in the previous examples. The results obtained are presented in Table IV below.

Example 5 - Flotation of Mixed Copper Sulfide Ore

Containing Molybdenum

A series of 30 gram samples of a -10 mesh

(U.S.) ore from Arizona containing a mixture of various copper oxide minerals and copper sulfide minerals plus minor amounts of molybdenum minerals was prepared. The grade of copper in the ore was 0.013 and the grade of the molybdenum was 0.00016. Each sample of ore was ground in a laboratory swing mill for 10 seconds and the resulting fines were transferred to a 300 ml flotation cell.

Each run was conducted at a natural ore slurry pH of 6.5. The collector (in the sodium salt form) was added at a dosage of 0.100 kg/ton of dry ore and the slurry was allowed to condition for 1 minute. Ore concentrate was collected by standard hand paddling between zero and 4 minutes. Just before flotation was initiated, a frother, a polyglycol ether available commercially from The Dow Chemical Company as Dowfroth^® 250 brand frother, was added in an amount equivalent to 0.030 kg/ton of dry ore.

The float cell in all runs was agitated at 1800 RPM and air was introduced at a rate of 2.7 liters per minute. Samples of the concentrates and the

tailings were then dried and analyzed as described in the previous examples. The results obtained are presented in Table V below.

Example 6 - Flotation of Iron Oxide Ore

A series of 600-g samples of iron oxide ore from Michigan was 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 collector was added and the slurry was allowed to condition for

1 minute. 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 was agitated at 900 RPM and air was introduced at a rate of 9.0 liters per minute.

Samples of the froth concentrate were collected at 4 minutes after the start of the air flow. Samples of the froth concentrate and the tailings were dried, weighed and pulverized for analysis. They were then dissolved in acid, and the iron content determined by the use of a D.C. Plasma Spectrometer. Using the assay data, the fractional recoveries and grades were calculated using standard mass balance formulas. The results are shown in Table VI below.

Claims

WHAT IS CLAIMED IS:

1. A process for the recovery of minerals by froth flotation wherein an aqueous slurry containing particulate minerals is subjected to froth flotation in the presence of mining reagents including at least one frother and at least one collector, the collector characterized in that it comprises a sulfonic component, and optionally other collector components, the sulfonic component consisting essentially of one or more sulfonic compounds selected from the group consisting of

(a) C_16-24 monoalkylated phenol sulfonic acids and salts thereof; (b) alkylaryl alkyl ether sulfonic acids or a salts thereof; and (c) alkylene oxide derivatives of monoalkylated phenol sulfonic acids or salts thereof.

2. The process of Claim 1 wherein the sulfonic component of the collector is selected from the group consisting of octadecyl anisole sulfonic acid, C_20-24 alkylated anisole sulfonic acid, C_24-28 alkylated anisole sulfonic acid, dodecyl anisole sulfonic acid, hexadecylphenyl butyl ether sulfonic acid,

octadecylphenyl butyl ether sulfonic acid, C_20-24 alkylphenyl butyl ether sulfonic acid, octadecyloxy cumene sulfonic acid, octadecyloxy toluene sulfonic acid, C_18-24 alkylatedphenyl isopropyl ether sulfonic acid, octadecylphenyl butyl ether sulfonic acid,

sulfonic acid of C_20-24 alkylphenol, sulfonic acid of octadecylphenol, sulfonic acid of hexadecylphenol, the salts thereof and mixtures thereof.

3. The process of Claim 1 wherein the sulfonic component of the collector is an alkylaryl alkyl ether sulfonic acid or salt thereof.

4. The process of Claim 1 wherein the sulfonic component of the collector is an alkylated phenol sulfonic acid or salt thereof.

5. The process of Claim 4 wherein the sulfonic component of the collector is selected from the group consisting of sulfonic acid of octadecylphenol, sulfonic acid of hexadecylphenol, salts thereof and mixtures thereof.

6. The process of Claim 1 wherein the sulfonic acid of the collector is an alkylene oxide derivative of an alkylated phenol sulfonic acid or salt thereof.

7. A process of Claim 1 characterized in that the sulfonic component of the collector corresponds to the following formula:

wherein X represents a covalent bond; ╌(CO)╌; or R³ wherein R³ is a linear or branched alkyl divalent moiety having 1 to 3 carbon atoms; R¹ is a C_1-30 linear, branched or cyclic alkyl group; R² is a C_1-30 linear, branched or cyclic alkyl group; and M is independently hydrogen, an alkali metal, alkaline earth metal, or ammonium or substituted ammonium, with the proviso that the total carbon content of R¹ and R² is from 16 to 34 carbon atoms.

8. The process of Claim 7 wherein R¹ is a

C_16-24 linear or branched alkyl group and R² is a C_1-4 linear or branched alkyl group.

9. The process of Claim 7 wherein R¹ is a C_16-24 linear or branched alkyl group and R² is

hydrogen.

10. The process of Claim 7 wherein R¹ is a C_16-24 linear or branched alkyl group and R² is

((-CH₂CHR⁴O)_aH wherein a is 1-2; and R⁴ is hydrogen, methyl or ethyl.