FI90398B - Collector of alkylated diaryloxide monosulphonates, which are usable in mineral flotation - Google Patents

Description

1 90398

This invention relates to the recovery of minerals by flotation.

Flotation is a process for treating finely divided mineral solids suspended in a liquid, for example powdered ore, which separates part of the solid from other fine mineral solids, for example quartz, siliceous side rock, clay or similar materials in the ore, a mass containing certain solids on the surface of a liquid in which the other solid components of the ore remain suspended (non-foaming). Flotation is based on the principle that the introduction of a gas into a liquid containing solid particles of various materials suspended causes some gas to adhere to certain suspended solids but not others, and makes the particles thus adhered to lighter than the liquid. Thus, these particles rise to the surface of the liquid as a foam.

The minerals and associated side rock that are handled by flotation are generally neither sufficiently hydrophobic nor hydrophilic to allow sufficient separation. Therefore, various chemical reagents are often used in flotation to create or improve the properties required for separation. Collectors are used to improve the hydrophobicity and thus the buoyancy of various valuable parts of minerals. Collectors must be able to (1) adhere to the desired mineral specimens and exclude other species present, (2) maintain dressing during flotation-related turbulence or surgery, and (3) make the desired mineral specimen sufficiently hydrophobic to the required degree of separation. allow.

In addition to collectors, a number of other chemicals are used. Examples of additional reagent types to be used include blowing agents, depressants, pH adjusting agents such as lime and sodium carbonate, dispersants, and various promoters and activators. Depressants are used to increase or enhance the hydrophilicity of various mineral washes and thus to reduce their foaming. Foaming agents are reagents that are added to flotation systems to promote the formation of semi-bilay foam. Unlike depressants and collector blowing agents, there is no need to adhere or adsorb to the mineral particles.

15 Flotation has been widely used in the mining industry since at least the beginning of the 20th century. Many different compounds are described as being useful as builders, blowing agents, and other blowing agents. For example, xanthates, simple alkylamines, alkyl sulfates, alkyl sulfonates, carboxylic acids, and fatty acids are generally accepted as useful builders. Reagents useful as blowing agents include lower molecular weight alcohols such as methyl isobutylcarbinol and glycol ethers. The additives used in the flotation 25 in each case are selected according to the nature of the ore, the flotation conditions and the mineral to be recovered, as well as the other additives to be used in the same context.

Although those skilled in the art will recognize that many different chemicals are useful in flotation, it is also known that the potency of known reagents will vary widely depending on the particular ore or ore being flocculated as well as the flotation conditions. It is further known that selectivity, i.e. the ability li 3 90398 to selectively foam the desired species and to exclude undesired species, is a particular problem.

Minerals and related ores are generally classified as sulfides or oxides, the latter group including oxygen-containing species such as carbonates, hydroxides, sulfates and silicates. Thus, minerals classified as oxides generally include all oxygenated minerals. Although much of the current minerals are in oxide ores, 10 a large proportion of successful flotation systems target sulfide ores. Flotation of oxide minerals is considered to be substantially more difficult than flotation of sulfide minerals, and the efficiency of most flotation methods in the recovery of oxide ores is limited.

15 One major problem associated with the recovery of minerals, both oxides and sulfides, is selectivity. Some of the known collectors, such as carboxylic acids, alkyl sulfates, and alkyl sulfonates mentioned above, are reported in the literature to be effective collectors in connection with oxide mineral ores. Although the use of these aggregators can lead to acceptable yields, it is known that the selectivity for the desired valuable mineral is typically quite poor. This means that the percentage of the desired component in the recovered mineral is lower than acceptable.

Due to the low concentration of separated mineral obtained by conventional direct flotation, the mining industry has generally switched to more complex methods in order to obtain minerals of acceptable concentration with acceptable intake. Oxide ores have often been sulfidized prior to flotation in existing industrial processes. Once the oxidide minerals have been sulfidized, they are flocculated using known sulfide collectors. Even with sulfidization steps, yields and concentrations are desirably less than 4,909,98. An alternative approach to the recovery of oxide ores is liquid-liquid extraction. A third approach used in the recovery of oxide ores, especially iron oxides and phosphates, is reverse or indirect flotation. Reverse flotation prevents foaming of an ore containing the desired valuable minerals and flotates side rock or other impurities. In some cases, an impurity is a mineral that may have value. A fourth approach to mineral recovery is chemical dissolution or extraction.

None of the existing oxide ore flotation methods is unproblematic. Generally speaking, known methods result in poor yield or concentration, or both. The low metal content of the recovered minerals is known to be a particular problem in the flotation of oxide minerals. Known recovery methods have not been economically viable, and therefore a large proportion of oxide ores simply remain untreated. Thus, those skilled in the art of flotation generally recognize the great need to improve selectivity in the flotation of oxidin minerals.

This invention relates to a process for recovering minerals by flotation, comprising flotation of an aqueous slurry containing particulate minerals in the presence of a collector which is diaryloxide sulfonic acid or a salt thereof or a mixture of such acids or salts wherein the monosulfonated species comprise at least about 20% sulfonated acids. -%, under conditions such that the minerals to be recovered 30 surface. The minerals to be recovered can be the desired minerals or undesirable impurities. In addition, the flotation method of the present invention uses flotation agents and other flotation reagents known in the art.

The use of the flotation process of this invention results in an improvement in selectivity and thus an improvement in the metal content of minerals recovered from oxide and / or sulfide ores while maintaining or increasing the total intake of the desired mineral. It is surprising that the use of alkylated diphenyloxide monosulfonic acids or their salts leads to uniform improvements in selectivity or the intake of valuable minerals.

The flotation process of this invention is useful in recovering valuable minerals from a variety of ores, including oxide ores as well as sulfide ores and alloys.

Non-limiting examples of oxide ores that can be foamed by the process of this invention are preferably iron oxides, nickel oxides, copper oxides, phosphorus oxides, aluminas and titanium oxides. Other types of oxygen-containing minerals that can be foamed by the process of this invention include carbonates such as calcite and dolomite, and hydroxides such as bauxite.

Non-limiting examples of oxide ores that can be recovered by flotation using the process of this invention include ores containing cassiterite, hematite, cuprite, vallerite, calcite, talc, kaolin, apatite, dolomite, bauxite, spinel, corundum, laterite, azurite, rutile, magnetite, columbite, ilmenite, smithsonite, anglesite-30, scheelite, chromite, serusite, pyrolusite, malachite, chrysococcus, zincite, bycite, massocot, rivet, gum mite, brucite, manganite, psilomelane, goethite, limonite, chrysoberyl, microlite, 35 tantalite, topaz and samarsite. The flotation process of this invention 6,90398 is useful in the treatment of other ores, including oxide ores, wherein the oxides are defined as carbonates, hydroxides, sulfates, and silicates, as well as oxides.

The process of this invention is also useful in the flotation of sulfide ores. Non-limiting examples of sulfide ores that can be foamed by the process of the invention include minerals containing chalcopyrite, chalcosite, lead compound, pyrite, zinc-10 mica, and pentlandite.

Precious metals such as gold and silver and platinum group metals including platinum, ruthenium, rhodium, palladium, osmium and iridium can also be recovered by the process of this invention. Such metals sometimes occur, for example, in connection with oxide and / or sulphide ores. Platinum, for example, sometimes occurs in association with troilite. By the process of this invention, such metals can be recovered in good yield.

20 Ores do not always exist as oxide or sulphide ores alone. Naturally occurring mammals can contain both sulfur- and oxygen-containing minerals as well as small amounts of the precious metals discussed above. By the process of this invention, 25 minerals can be recovered from these alloy ores. This can be done by two-stage flotation, where the second stage is a conventional sulfide flotation that recovers predominantly sulfide minerals, and the second flotation stage uses the method and collector composition of this invention to recover predominantly oxide minerals and any noble metals present. Alternatively, both sulfur-containing and oxygen-containing minerals can be recovered simultaneously by the process of this invention.

I; 790398 One particular feature of the process of this invention is the ability to foam different minerals in different ways. Without wishing to be bound by any theory, it is thought that the susceptibility of various minerals to foam in the process of the invention is related to the crystal structure of the minerals. More specifically, it appears that there is a correlation related to the relationship between the lengths of the crystal edges and the surface area of the crystal in terms of unit area. Minerals with higher ratios appear to rise to the surface primarily compared to minerals with lower ratios. Thus, minerals with at least 24 surfaces in the crystal structure (Group I) are generally more likely to rise to the surface than minerals with 16 to 24 surfaces (Group 15 II). Group III minerals with 12 to 16 surfaces are next in ascending order, followed by Group IV minerals with 8 to 12 surfaces.

In the process of the present invention, Group I minerals generally rise to the surface before Group II minerals, which rise before Group III minerals, which in turn rise to the surface before Group IV minerals. By prior, i.e., primary rise to the surface, is meant that the preferred species rises to the surface at a lower dosage (amount) of collector used. This means that the group I mineral can be recovered at a very low dosage. After increasing the dosage and / or removing most of the group I mineral, the group II mineral, etc. is recovered.

30 One skilled in the art will appreciate that these groupings are not absolute. Different minerals may have different possible crystal structures. In addition, the size of naturally occurring crystals varies, which affects how easily different minerals can be foamed. An additional factor that affects the order of foaming is the purity of 905S8 substitution. In addition, within a group, i.e., among minerals with similar edge length to surface area ratios, these and other factors affect which of the group members first rises to the surface.

5 It is easy for a person skilled in the art to determine to which group a mineral belongs by examining the usual mineralogical descriptions of the various minerals. They are described, for example, in Cornelius S. Hurlbut, Jr. and Cornells Klein, Manual of Mineralogy, 19th ed., John Wiley and Sons, New York 1977. 10 Non-limiting examples of Group I minerals include graphite, nickelite, covellite, molybdenum and beryl.

Non-limiting examples of Group II minerals include rutile, pyrolucite, cassiterite, anatase, calomel, trobernite, autunite, maralite, meionite, apophilite, zirconium and xenotime.

Non-limiting examples of Group III minerals include arsenic, greenockite, millerite, zincite, corundum, hematite, brucite, calcite, magnesite, siderite, rhodocrosite, smithsonite, Chilean alpha, apatite, pyromorphite, mimetic and vanadium.

Non-limiting examples of Group IV minerals include sulfur, chalcosite, chalcopyrite, stib-25 rivet, bismuthinite, löllingite, marcasite, mass-bag, brookite, bohmite, diaspori, gothite, samars-kite, attackamite, aragonite, vitreous, vitreous, , phosgenite, potassium tartar, thenarite, barite, celestite, anglesite, anhydrite, 30 epsomite, antlerite, chaldonite, trifylite, lithiophilite, heterosite, purpurite, variscite, strengite, chrysoberyl, scorodite, scorodite, libetenite, adamite, phosphuranylite, childrenite, eosphorite, h 9 90398 scelite, powellite, wulfenite, topaz, columbite and tantalite.

As mentioned above, these groupings are theoretically useful in determining which minerals 5 rise more easily to the surface. However, the collector and method of this invention are useful in foaming various minerals that do not fit the above groups. These groupings are useful in predicting which minerals rise to the surface with the lowest relative collector partition, but not in determining which minerals can be recovered by the flotation method of the invention.

The selectivity of the collectors of this invention makes it possible to separate small amounts of undesired minerals from the desired minerals. For example, the presence of apatite is often a problem in iron flotation. Thus, the collectors of this invention are in some cases useful in reverse flotation, where an undesired mineral is raised to a surface, such as by separating topaz from foamed cassiterite or apatite from iron.

In addition to flotation of naturally occurring ores, the flotation method and collector composition of this invention are useful in flotation of minerals from other sources. One such example is waste materials from various processes such as heavy medium separation, magnetic separation, metalworking and oil refining. These materials often contain minerals that can be recovered using the flotation method of this invention. Another example is the recovery of a mixture of graphite ink and other carbon-based inks during paper recycling. Such recyclable papers are typically deinked to separate the printing inks from the paper-35 fibers by a flotation process. The flotation method of the present invention is particularly effective in such flotation flotation methods.

The diaryloxide monosulfonic acid or monosulfonate collector of this invention corresponds to general formula 5

Ar'-O-Ar, wherein each of Ar 'and Ar is independently at each occurrence a substituted or unsubstituted aromatic group, such as a phenyl or naphthyl group, provided that one and only one of Ar' and Ar contains one sulfonic acid or sulfonic acid salt group. The diaryloxide monosulfonic acid or monosulfonate collector is preferably an alkylated diphenyl oxide or alkylated biphenylphenyl oxide monosulfonic acid or monosulfonate or a mixture thereof. The diaryloxide monosulfonic acid or monosulfonate is preferably substituted with one or more hydrocarbyl substituents. Hydrocarbyl substituents may be substituted or unsubstituted alkyl groups or substituted or unsubstituted unsaturated alkyl groups.

More preferably, the monosulfonated diaryloxide scavenger of this invention is a diphenyloxide scavenger and corresponds to the following formula or is a mixture of compounds corresponding to the following formula: (R) m (R) n cp ~ ”-cp (SO 3 M +) y (SO 3 M +) x 35 li 11 90398 wherein each of R is independently a saturated alkyl group, a substituted saturated alkyl group, an unsaturated alkyl group or a substituted unsaturated alkyl group; each of m and n is independently 0, 5 1 or 2; each group M is independently hydrogen, an alkali metal, an alkaline earth metal, or an ammonium or substituted ammonium group, and each of x and y is independently 0 or 1, provided that the sum of x and y is one. Preferably, each R group of one or more is independently an alkyl group having 1 to 24, more preferably 6 to 24, even more preferably 6 to 16, and most preferably 10 to 16 carbon atoms. Alkyl groups may be linear, branched or cyclic groups, of which linear or branched groups are preferred. It is also preferred that each of the numbers m and n is one. Ammonium ion groups M * have the formula (R ') 3HN', wherein each of R 'is independently a hydrogen atom or a C 1-4 alkyl or O, -4-hydroxyalkyl group. Illustrative examples of C 1-4 alkyl and hydroxyalkyl groups include methyl, ethyl, propyl, isopropyl, butyl, hydroxymethyl and hydroxyethyl groups. Typical ammonium ion groups include ammonium (N * H4), methylammonium (CH3N * H3), ethylammonium (C2H5N * H3), dimethylammonium [(CH3) 2N * H2], methylethylammonium (CH3NtH2C2H5), trimethylammonium [* H], di-methylbutylammonium [(CH3) 2N * HC4H9], hydroxyethylammonium (HOCH2CH2N * H3) and methylhydroxyethylammonium (CH3N * H2CH2CH2OH). Each of the groups M is preferably hydrogen, sodium, calcium, potassium or ammonium.

Alkylated diphenyloxide sulfonates and methods for their preparation are well known and are referred to in the context of this invention. The monosulfonate collectors of this invention can be prepared by modifications of known methods for preparing sulfonates. Representative examples of methods for preparing sulfonates are disclosed in U.S. Patent Nos. 3,264,242, 3,634,272, and 3,945,437.

i2 90 398 242; Commercial methods for preparing alkylated diphenyloxide sulfonates generally do not result in species that are exclusively monoalkylated, monosulfonated, dialkylated, or sulfonated. Commercially available species are predominantly (greater than 90%) mixtures of disulfonated and mono- and dialkylated species with 15-25% dialkylated species and 75-85% monoalkylated species. Commercially available species are typically about 80% monoalkylated and 20% dialkylated.

In connection with the practice of this invention, the use of a monosulfonated species has been found to be crucial. Such monosulfonated species can be used by modifying the sulfonation step in methods described, for example, in U.S. Patent Nos. 3,264,242, 3,634,272, and 3,945,437. More specifically, the methods described above relate primarily to the preparation of disulfonated species. Thus, in the sulfonation step, it is advised to use sufficient sulfonating agent to sulfonate both aromatic rings. However, in preparing the monosulfonates useful in the practice of this invention, it is preferred to limit the amount of sulfonating agent used to that required to provide one sulfonate group per molecule.

Monosulfonates prepared in this manner contain both molecules that are not sulfonated and molecules that contain more than one sulfonate group per molecule. The monosulfonates can be separated and used in relatively pure form, if desired. However, the mixture obtained in the sulfonation step using only an amount of sulfonating agent that provides approximately one sulfonate group per molecule is also useful in the practice of this invention.

Il i3 90 398

As mentioned above, the use of monosulfonated species is critical to the practice of this invention. However, the presence of disulfonated species is not theoretically considered harmful in the presence of at least 5 to 20% of monosulfonated species. It is preferred that the proportion of monosulfonated species is at least 25%, more preferably at least 40% and most preferably at least 50%. It is most preferred to use relatively pure monosulfonated acids or salts. In the context of commercial applications, one skilled in the art will appreciate that the higher costs associated with the preparation of relatively pure monosulfonated species are offset by the loss of performance associated with the use of mixtures containing disulfonated species.

Commercially available alkylated diphenyl oxide sulfonates are often mixtures of monoalkylated and dialkylated species. Although such mixtures of monoalkylated and dialkylated species are useful in the context of this invention, in some circumstances it is preferred to use species that are either monoalkylated, dialkylated or trialkylated. Such species are prepared, for example, by modifications of the methods of U.S. Patent Nos. 3,264,242, 3,634,272, and 3,945,437. When it is desirable to use a mixture other than mixture, after the alkylation, a distillation step is performed to remove the monoalkylated species and monoalkylated species are used or recycled for further alkylation. In general, it is preferred to use dialkylated species, although monoalkylated and trialkylated species are also functional.

Non-limiting examples of preferred alkylated diphenyloxide sulfonates include monosulfonated sodium diphenyloxide, monosulfonated sodium hexyldiphenyloxide, monosulfonated sodium decyldiphenylphenoxyphenyldisodium diphenyloxy-di-diphenyldiphenyloxy In a more preferred embodiment, the collector is monosulfonated sodium dialkyldiphenyloxide, wherein the alkyl group is a C10-16 alkyl group, most preferably a C10-12 alkyl group. Alkyl groups can be branched or straight chain.

The collector can be used at any concentration that provides the desired selectivity and intake of the desired valuable minerals. The concentration to be used depends in particular on the mineral to be recovered in each case, the mineral content of the ore to be foamed and the desired quality of the mineral to be recovered.

Additional factors to be considered in determining dosages include the surface area of the ore to be treated. As will be appreciated by those skilled in the art, the smaller the particle size, the larger the surface area and the greater the amount of collector required to provide adequate uptake and concentration. Oxide mineral ores typically need to be ground finer than sulfide ores, and therefore require very large 20 batches or removal of the finest particles by sludge removal. Conventional oxidizing mineral flotation methods typically require a sludge removal step to remove the fines present and thus allow the method to operate at acceptable collector dosages. The collector of this invention operates at an acceptable dosage with or without sludge removal.

The collector content is preferably at least 0.001 kg / t, more preferably at least 0.05 kg / t. It is also preferred that the total concentration of the collector is at most 5.0 kg / t, more preferably at most 2.5 kg / t. In order to achieve optimal performance of the collector, it is generally most advantageous to start with low doses and increase the dosage until the desired effect is achieved. Although the improvement in yield and concentration achieved by the present invention increases with increasing 90399 dosage, those skilled in the art will appreciate that at some point, the improvement in intake and concentration provided by the higher dosage will be offset by increased foaming chemical costs. It will also be appreciated by those skilled in the art that varying aggregate doses are required depending on the type of ore and other flotation conditions. In addition, it has been found that the required collection dose is proportional to the amount of mineral to be recovered. In situations where the amount of mineral foamable by the method of the present invention is small, a very small amount of collector is required due to the selectivity of the collector.

In the recovery of certain minerals, it has been found advantageous to add a collector to the flotation system in stages. By stepwise addition is meant that a portion of the accumulator dosage is added; recovering the foamed concentrate; a new batch of collector is added and the foamed concentrate is recovered again. It is not advantageous to change the total amount of collector used when adding the collector step by step. This stepwise addition can be repeated a few times to achieve optimal intake and concentration. The number of steps to add a collector is limited only by practical and economic factors. Preferably, up to about six steps are used.

An additional advantage of gradual addition relates to the ability of the collector of this invention to foam different minerals differently as the dosage changes. As mentioned above, at low dosages, only the mineral which is particularly easy to foam with the collector according to the invention rises to the surface, while the other minerals remain in the sludge. As the dose increases, a different mineral may rise to the surface, allowing the different minerals in a particular ore to be separated.

In addition to the collector of this invention, other conventional reagents or additives may be used in the flotation. Examples of such additives ie 90398 include various depressants well known to those skilled in the art! t and dispersants. In addition, it has been found useful to use compounds containing hydroxyl groups, such as alkanolamines or alkylene glycols, to improve selectivity for the desired minerals in systems containing resin or siliceous side rock. The collector of this invention may also be used in conjunction with other collectors. In addition, blowing agents can and typically are used. Foaming agents are well known in the art and are referred to in the context of this invention. Examples of useful blowing agents include polyglycol ethers and relatively low molecular weight foaming alcohols.

A particular additional advantage of the collector according to the present invention is that no other additives are needed to adjust the pH of the flotation slurry. The flotation process using the collector of this invention operates efficiently in the typical inherent pH range of ores from 5 to 20 lower to 9. This is particularly important given the reagent costs required to adjust the pH of the slurry from a natural value of up to about 7. 9.0 to 10.0 or above, which is typically necessary when using conventional carboxyl, sulfone, phosphone and xanthine scavengers.

The ability of the collector of this invention to operate at a relatively low pH means that it can also be used if it is desired to lower the pH of the slurry. The lower limit for the pH at which the collector of this invention operates is the pH at which the surface charge of the mineral nest is suitable for collector adhesion.

Since the collector of the present invention operates at different pH values, it is possible to take advantage of the tendency of different minerals to rise to the surface at different pH values. This makes it possible to make one flotation of li i7 90398 at a certain pH to optimize the flotation of a particular species. The pH can then be adjusted to the next flotation to optimize foaming of the different species, thereby facilitating the separation of the various minerals present together.

The collector of this invention can also be used in conjunction with conventional collectors. For example, the monosulfonated diaryloxide collectors of this invention can be used in a two-stage flotation in which monosulfonated diaryloxide flotation captures primarily oxide minerals, while a second flotation step using conventional collectors is used primarily to recover additional sulfide minerals or oxide minerals. In the case of conventional collectors, a two-stage flotation can be used, in which the first stage comprises the procedure according to the invention and is carried out at the natural pH of the slurry. The second step uses conventional collectors and is carried out at an elevated pH. It should be noted that in some circumstances it may be desirable to reverse the order of these steps. The advantage of such a two-step process is the consumption of a smaller amount of additive to adjust the pH, and it also allows a more complete recovery of the desired minerals by performing flotation under different conditions.

The following examples are provided to illustrate the invention and should not be construed as limiting it in any way. Unless otherwise stated, all parts and percentages are by weight.

30 The following examples include flotation experiments performed in a Hallimond tube and laboratory-sized flotation cells. It should be noted that Hallimond tube flotation is a simple way to screen collectors, but it does not necessarily predict collector success in actual flotation. Hallimond Pipe Foaming 90398 does not use shear or mixing in actual flotation and does not measure the effect of flotation agents. Thus, while a collector must be effective in Hallimond tube flotation if it is to be effective in actual flotation, in an Hallimond tube an efficient collector may not be effective in actual flotation. It should also be noted that experience has shown that the collector doses required to achieve satisfactory yields in a Hallimond tube are often substantially higher than those required in the flotation cell test. Thus, Hallimond tube experiments cannot accurately predict the doses that would be required in a true flotation cell.

Example 1 15 Hallimond pipe foaming of malachite and quartz

Approximately 1.1 g of (1) malachite, a copper oxide mineral of the approximate formula Cu 2 CO 3 (OH) 2, or (2) quartz was screened for approximately -60 to +120 US mesh and placed in a small flask with deionized water 20 (about 20 mL ). The mixture was shaken for 30 s and then the aqueous phase containing some unsuspended fines, i.e. sludge, was decanted. This sludge removal step was repeated a few times.

150 ml of deionized water was placed in a 250 ml de-25 beaker. Next, 2.0 ml of 0.10 M potassium nitrate solution was added as a buffer electrolyte. The pH was adjusted to approximately 10.0 by the addition of 0.10 N HCl and / or 0.10 N NaOH. Next, 1.0 g of demineralized mineral was added, along with an amount of deionized water 30 to give a total volume of about 180 ml. Collector, branched C16 alkyl substituted sodium diphenyloxide sulfonate with a monoalkylated species content of about 80% and a dialkylated species content of about 20% was added and stirred for 35 minutes. The pH was monitored and adjusted as necessary

II

19 90 398 with HCl and NaOH. It should be noted that Experiments 1-5 are not embodiments of the invention and use a disulfonated collector, while Experiments 6-10, which are embodiments of the invention, use a monosulfonated collector. The only difference is that some are disulfonated and others are monosulfonated.

The slurry was transferred to a Hallimond tube designed to attach a hollow needle to the bottom of a 180 ml tube. After the slurry was added to the Hallimond-10 tube, the tube orifice was connected to a reduced pressure of 16.9 kPa for 10 min. This vacuum allowed air bubbles to enter the tube through a hollow needle placed at its bottom. The slurry was stirred during flotation with a magnetic stirrer at 200 rpm.

15 The foamed and non-foamed material was separated from the slurry by filtration and dried in an oven at 100 ° C. Each part was weighed; the yields of copper and quartz are shown in the following Table I. After each experiment, all equipment was washed with concentrated HCl and rinsed with 0.10 N NaOH solution and doionized water before the next experiment.

The reported copper and quartz yield is the proportion of mineral recovered from the original mineral placed in the Hallimond tube. Thus, a yield of 1.00 indicates that 25 material has been completely recovered. It should be noted that although copper and quartz yields are reported together, the results have actually been obtained in two experiments performed under similar conditions. It should further be noted that low quartz yield indicates selectivity for copper 30. The accuracy of the copper yield values is generally ± 0.05 and that of the quartz yield values is generally ± 0.03.

2o 90 3 98

table 1

Dosage for Cu- For quartz- 5 Collector (kg / kg) pH ti proportion ti proportion 1® LC, 6DPO (SOJNa) 2 <3> o, o60 5.5 0.760 0., 153 2® LC, 6DP0 ( 50jNa) 2® 0.060 7.0 0.809 0.082 3® LC, 6DPO (SO, Na) 2® 0.060 8.5 0.800 0.062 »I® LC, 6DPO (SOjNa) 2® 0.060 10.0 0.546 0.104 10 5® LC, 6DP0 (SO3Na) 2® 0.060 1 1.5 0.541 0.130 ~~ 6 LC, 6DPO (SO3Na), ® 0; 060 5.5 0.954 0.135 7 LC, 6DPO (SO} Na), ® 0.060 7.0 0.968 0.097 8 LC, 6DPO (SC> 3Na), ® 0.060 8.5 0.913 0.084 15 9 LC (6DPO (SOjNa) i® 0.060 10.0 0.837 0.070 10 L CeDPOiSOjNa), ® 0.060 1 1.5 0.798 0.065 1. For linear C16- alkyl-substituted diphenyloxide sulfonate containing about 80% ion-20 and 20% dialkylated species and sold by The Dow

Chemical Company as a surfactant DOWFAX ™ 8390 2. Embodiment not according to the invention 3. Diphenyloxide monosulphate substituted by linear Cu-alkyl groups containing about 80% of mono- and 20% of dialkylated species

The results shown in Table I above clearly demonstrate the efficiency of the collectors of this invention. A comparison of Experiments 1-5, which are not embodiments of the invention 30, and Experiments 6-10 shows that the monosulfonated collector of this invention consistently resulted in substantially higher copper yields and comparable or lower quartz yields at different pH values.

Il 2i 90398

Example 2

Foaming of iron oxide ore

A series of 600 g samples of iron ore from Michigan was prepared. The ore contained a mixture of hematite, mar-5 tite, goethite, and magnetite mineral scrubbers. Each 600 g sample was ground with deionized water (400 g) in a hammer mill for 10 min at approximately 60 rpm. The resulting mass was transferred to 3,000 ml

To an Agitair flotation cell equipped with an automatic 10 blade removal system. A collector was added and the slurry was allowed to stand for 1 min. Next, a polyglycol ether blowing agent in an amount of 40 g / t of dry ore was added and allowed to stand for another 1 min.

The flotation cell was agitated at 900 rpm for 15 min and air was fed at 9.0 l / min. The foamed concentrate was sampled 1.0 and 6.0 min after the start of the air supply. The foam concentrate and residual samples were dried, weighed and ground for analysis. They were dissolved in acid and the iron content was determined using direct current plasma spectrometry.

Yield fractions and concentrations were calculated from the assay results using standard mass equilibrium formulas. The results are shown in Table II below.

22 90 398 tn P · Η

3 ^ _ -P

ui o - = r 4j • H in VO rd o - = r jt · η p

Jp -P t. Rv -pc tn -h Q (-s -p o C Oj ^ u p tn 0 P r — l

0) CP

-p o w s: vw o

X 2 t fH G

c o o co P vO 0 * n tn £

P · *> r * H * H

en O O 3 2 en tn

tn tn χ X

3 3 O O

3 D -H -H

M CO “Τ” 'r — t i — C

• H -H CTv O> i>.

0 e o en. = R> i> <

Ί-1 Ή -P g ~ CC

• 3 E -H _ CL) <U

& DJ O O Μ-l m 1 tO * 3 -3 Ό T3 oi e e n -H 3 3 rH 4-1 VO O · 3 -3

• H C O = T 1-1 U

<Q <- (\ i + J -P

G 3 o »> 3 3

M 3 in CC

m 3 o a „2 3 3

O 3 tn 4-i P

Ai 4-> 3 -3-3

Ai 3 3 äi 0 0 3 3 en <M t— 3 3

Ή CC -H vO OO P P

3 O 3- ~ r * 3 -h 3 C P ^ ^ p p E- · -h -h _ _ tn en e til0 ° λ λ

P P

t-η tn en • H · Η

1 np V

h f—

o Ό α \ ί— O: p * P

C «sy vO -P fH <H

P rv r * 0 Ή »H

P r-% <- «» P * H -H

tn ° ° see tn s: x: P> 1> 1 -p u u

H · Η -H

tO __ P cH rH

3 0 0 O> 1>, 3 ^ 0 0 3> 1> 1 tn p> oa oj tn x x

O \ w *> <H rH

G tT * λ Cp P

C M ° ° (Dl I

<^ C cv cv

H rH rH

WOOD

λ λ P p p

® »e rH H

^ ^ »H f-H

rt) C -H -H

2 2 sO o φ

p *> rn CC C

O O e p p

P -H -p 4J

• n 5-- w tn -p + j O O O · Η · Η o o. o. 000 ^ D Q * p p

o p P P P

U * H p p; q q ω n: n ““ H c ^ ^ <i> I Λ o Θ

x, \ * - C \ J

I! 23 90398

A comparison of Experiments 1 and 2 shows that the use of a monosulfonated collector according to the present invention resulted in about 50% better yield and a slightly higher concentration of iron than the use of a disulfonated collector.

5 Example 3

Flotation of rutile ores

A series of samples (30 g) were prepared from a mixture having a particle size of -10 US mesh and containing 10% rutile (TiO 2) and 90% quartz (SiO 2). Each ore sample 10 was ground with deionized water (15 g) in a hammer mill (diameter 6.35 cm, rod size 1.27 cm) for 240 spins. The resulting mass was transferred to a 300 ml flotation cell.

The pH of the slurry was allowed to be at a natural ore value of 8.0. After the addition of the collector according to Table III, the slurry was allowed to stand for 1 min. Next, a blowing agent, polyglycol ether, corresponding to 0.050 kg / t of dry ore was added, and the slurry was allowed to stand for another 1 min.

The flotation cell was agitated at 1,800 rpm and air was fed at 2.7 l / min. Foam concentrate samples were taken on a standard hand pallet at 1.0 and 6.0 min after the start of the air supply. The concentrate and residual samples were dried and analyzed before use. 25 The results obtained are shown in Table III below.

24 90398 w a t— \ r \ = r d σο o cr> t— oo σ 'tn o * - ο o o o: <o -Η Γ ^ ^ «'« - £ ° oooooo

01 -H

¢) qJ

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co oi OOOOOO

• H CO

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03 t * H f— ΓΟ ON q- H ^ r-votnO'-C'-

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C/O

0 OOOOOO

^ -oooooo c σ> ^ -> -v «c ^ oooooo <- © ® ® © ® ® CN» - Γ- (* - ♦ - «- m <9 nj (D <n to zzzzzzmmmm rn oooooo νΛ * S > ΙΛ \ f \ \ / \ reason oooooo 9? CL Q- cl o. Cl a.

2 a a o o o o

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* o Q Q O o Q

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II

25 90 398

Table III (continued) 1. No embodiment of the invention 2. disubstituted by a linear C16 alkyl natriumdifenyylioksididisulfonaatti 5 3. Linear di-substituted with C16 alkyl natriumdifenyylioksidimonosulfonaatti 4. disubstituted branched-chain C12-alkyl natriumdifenyylioksididisulfonaatti 5. A branched C12 alkyl disubstituted 10 tu 6. The branched natriumdifenyylioksidimonosulfonaatti Sodium diphenyloxide monosulfonate disubstituted with C10 alkyl groups

The results shown in Table III above demonstrate the effectiveness of the collector of this invention in improving titanium yield and content. Comparison of Experiment 1 to Experiment 2 and Comparison of Experiments 4 and 5 to Experiment 3 show the significant improvements provided by the use of the monosulfonate collectors of this invention compared to disulfonate-20 collectors.

Example 4

Separation of apatite and quartz

A series of samples (30 g) with a particle size of -10 US mesh and containing 10% apa-25 titite [(Ca5 (ltF) (PO4) 3] and 90% quartz (SiO2) were prepared. similar to that used in Example 3. The inherent pH of the ore was 7.1 A mixture of monosulfonated and disulfonated collector was used in Experiments 8 to 13. The results shown in Table IV demonstrate the ability of the process of this invention to distinguish between apatite and quartz.

26 90398

Table IV

Dosage

Collector iiig / t) p-yield p-content 1® L, DC, 0DPO (SO3Na) 2® 0.050 0.115 0.08 1 2 L, DC, 0DPO (SO3Na), ® 0.050 0.962 0.068 3® 8, D-C12DP0 (SO3Na) 2® 0.050 0.235 0.078 4 B, D-C12DP0 (SO3Na) i® 0.050 0.989 0.067 5 Refined kerosene® 0.050 LD-C10DPO (SO3Na) i® 0.050 0.925 0.103 6 Refined kerosene® 0.010 L, DC, 0DPO (SO3Na ), ® 0.050 0.862 0.112 7 Refined kerosene® 0.020 L, D-C10DPO (SO3Na) i ® 0.050 0.818 0.125 8 L, DC, 0DPO {SO3Na) 2® 0.090 L, DC, 0DPO (SO3Na) 1® 0.010 0.336 0.077 9 L, D-ClODPO (5O3Na) 2® 0.030 L, DC, ODPO (SO3Na), ® 0.020 0.529 0.075 10 L, DC, ODPO (SO3Na) 2 © 0.020

LrD-C, ODPO (SO3Na), ® 0.030 0.699 0.079 11 L, D-C10DPO (SO3Na) 2® 0.010 L, DC, ODPO (SO3Na), ® 0.090 0.866 0.069 12 LrD-C, ODPO (SO3Na) 2® 0.080 L, D-C10DPO (SO3Na), ® 0.020 0.539 0.067 13 LD-C, ODPO (SO3Na) 2® 0.160

LrD-C, ODPO (SO3Na), ® 0.090 0.877 0.053 27 90 398

Table IV (continued) 1. In the embodiment according to the invention, the El 2. Linear di-substituted with C10 alkyl natriumdifenyylioksididisulfonaatti 5 3. Linear di-substituted with C10 alkyl natriumdifenyylioksidimonosulfonaatti 4. disubstituted branched-chain C12-alkyl natriumdifenyylioksididisulfonaatti 5. A branched C12 alkyl disubstituoi TU-10 natriumdifenyylioksidimonosulfonaatti 6. Processed kerosene product sold by Philips Petroleum as kerosene Soltrol ™. It is added to the flotation cell at the same time as the collector.

15 The data presented in Table IV show the considerable efficiency of monosulfonated collectors in the recovery of phosphorus from ores containing apatite and quartz. Comparison of Experiments 2-4 to Experiments 1 and 2, which were not examples of the invention, shows the effect of monosulfonation. Experiments 5 and 6 show that the collector of this invention was effective when used in conjunction with added hydrocarbon. The slight decrease in yield was accompanied by a significant increase in concentration. Experiments 8-13 show the effect of mixing monosulfonated collectors and disulfonated co-collectors. Comparison of Experiments 2, 11, and 13, in which the concentrations of monosulfonated collectors are comparable and the concentration of the disulfonated species ranges from zero to 0.160 kg / t, shows that the disulfonated species appears to act as a diluent in the presence of low concentrations. At higher concentrations, disulfonated species do not interfere with intake, but appear to lower metal content.

Example 5

Samples (30 g, -10 US mesh) of ore from Kes-35 Ki-Africa were prepared. Approximately 90% of the copper metal content of the ore 2β 90398 was in malachite and the remainder in other copper minerals. Each ore sample was ground with deionized water (15 g) in a miniature rod mill (diameter 0.35, rods 1.27) for 1,200 revolutions. The resulting mass 5 was transferred to a 300 ml miniature flotation cell. The pH of the slurry was allowed for Malmi to have a natural value of 6.2. The collector was added 0.250 kg / t dry ore in experiments 1-20. In Experiments 20-26, the collector dosage was varied and in Experiments 22-26, the collector contained varying amounts of di-10 sulfonate. After the addition of the collector, the slurry was allowed to stand in the cell for 1 min. Next, a blowing agent, polyglycol ether, was added at a rate of 0.080 kg / l dry ore. This addition was followed by a 1 min stand.

The flotation cell was agitated at 1,800 rpm and air was fed at 2.7 rpm. Foam dew was collected for 6.0 min. Concentrate and residue samples were dried, weighed, ground for analysis and dissolved in acid. The copper content was determined by direct current plasma spectrometry.

Il 29 90 398

Table V

Dosage Cu- Cu-

Test Collector (kg / h) pH intake concentration 1 © no collector - 6.2 0.038 0.019 2 BD-C ^ DPCKSOjNah® 0.250 6.2 0.696 0.057 3® B, D-C12DPO (SO3Na) 2® 0.250 6.2 0.501 0.042 4 L, D-C10DPO (SO3Na), © 0.250 6.2 0.674 0.056 5® L, DC, oDPO (SO3Na) 2® 0.250 6.2 0., 487 0.035 6 L, DC, 0BIPPE (SO3Na), ® 0.250 6.2 0.696 0.059 7® L, DC, 0BIPPE (SO3Na) 2® 0.250 6.2 0.573 0.051 8 L, DC, 6DP0 (SO3Na), ® 0.250 6.2 0.714 0.058 9® L, D-C16DPO (SO3Na) ) 2® 0.250 6.2 0.598 0.052 10 L, M-C10DPO (SO3Na), 10 0.250 6.2 0.390 0.046 11® L, M-ClODPO (SO3Na) 2H 0.250 6.2 0.1 16 0.038 12 B, M -Cl2DPO (SO3Na), 12 0.250 6.2 0.338 0.044 13® B, M-C12DPO (SO3Na) 2 »3 0.250 6.2 0.145 0.041 14 L, M-C24DPO (SO3Na), 14 0.250 6.2 0.474 0.037 15 ® L, M-C24DP0 (503Na) 215 0.250 6.2 0.335 0.035 16 LM-QDPCKSOsNahie 0.250 6.2 0.11 1 0.037 17® L, M-C60PO (SO3Na) 217 0.250 6.2 0.053 0.038 18 L, D -C6DPO (SO3Na), 18 0.250 6.2 0.317 0.04 1 19® LD-C6DP0 (SO3Na) 219 0.250 6.2 0.198 0.038 30 90 398

Table V (continued) 5

Dosage Cu- Cu-

Experiment Collector (kg / h) pH yield concentration 20 BD-C ^ DPCHSOaNa), © 0.400 6 ^ 2 0, B39 ~ 0.055 10 21 © B (DC, 2DP0 (SO3Na) 2 © 0.400 6.2 0.533 0.039 22 B, D-Cl2DP0 (503Na), ® 0, 100 6.2 0.620 0.045 B, D-C12DP0 <SO3Na) 2 © 0.300 ____ 23 B, D-C12DP0 (SO3Na), ® 0.200 6.2 0.683 0.051 B, D-C12DP0 (SO3Na) 2® 0.200 ______ 24 B, D-C12DPO (503Na), @ 0.300 6.2 0.788 0.054 ib B, DC, 2DP0 (SO3Na) 2® 0.100 25 B, D-C12DPO (SO3Na), @ 0.400 6, 2 0.855 0.041 B, DC, 2DP0 (SO 3 Na) 2® 0.400___ 26 B (DC, 2DP0 (SO 3 Na), ® 0.400 6.2 0.861 0.039 BD-C, 2DPO (SO 3 Na) 2® 1, 200 20 1. El of the invention 2. In the embodiment according to branched C12-disubstituted with alkyl groups natriumdifenyylioksidimonosulfonaatti 25 3. A branched C12 alkyl di natriumdifenyylioksididisulfonaatti 4. Linear di-substituted with C10 alkyl natriumdifenyylioksidimonosulfonaatti 5. Linear di-C 10 -C 30 alkyl natriumdifenyylioksididisulfonaatti 6. Linear di-substituted with C10 alkyl bifenyyliee tterimonosulfonaatti 7. Linear di-substituted with C10 alkyl difenyylieetteridisulfonaatti 3i 90 398 8. Linear di-substituted with C16 alkyl natriumdifenyylioksidimonosulfonaatti 9. Linear C16 disubstituted with alkyl groups natriumdifenyylioksididisulfonaatti 5 ~ 10 linear C10 alkyl monosubstituted natriumdifenyylioksidimonosulfonaatti 11. The linear C10-alkyl-substituted-monosubsti natriumdifenyylioksididisulfonaatti 12. The branched C12 alkyl group mono-10-substituted natriumdifenyylioksidimonosulfonaatti 13 branched C12-alkyl-substituted natriumdifenyylioksididisulfonaatti monosubsti-14 linear C24 alkyl-substituted monosubsti-natriumdifenyylioksidimonosulfonaatti 15 15 C 24 linear alkyl group monosubstituted natriumdifenyylioksididisulfonaatti 16. The linear C6 alkyl group monosubstituoi-tu natriumdifenyylioksidimonosulfonaatti 17. With linear C6 alkyl monosubstituted sodium diphenyloxide disulfonate 18. Sodium diphenyloxide monosulfonate disubstituted with linear C6 alkyl groups 19. Sodium diphenyloxide disulfonate disubstituted with linear C6 alkyl groups. . : 25 --- The data presented in the table above show the efficiency of various alkylated diaryloxide monosulfonates in the flotation of copper oxide ores. A comparison of even-numbered experiments 2-18, which are examples according to the invention 30, and odd-numbered experiments 1-19, which are also not according to the invention, clearly shows substantially improved results obtained using a monosulfonated collector compared to using a disulfonated collector at the same concentration. tena. Comparison of Experiment 2 with Experiment 21 shows its effect on dosage. Experiments 20-26 show that the disulfonated species in the mixtures appear to act as a diluent when mixed with the monosulfonated collectors of this invention.

5 Example 6

Foaming of iron oxide ore

A series of samples (600 g) of iron oxide ore from Michigan were prepared. The ore contained a mixture of hematite, martite, goethite, and magnetite mineral-10 specimens. Each 600 g sample was milled with deionized water (400 g) in a cloth mill at approximately 60 rpm for 15 min. The resulting mass was transferred to a 3,000 mL Agitair flotation cell equipped with an automatic paddle removal system. at a natural pH of 7.0 of the slurry.

The amount of propylene glycol mentioned in Table VI below was added and the slurry was allowed to stand for 1 min. Next, a collector was added and the slurry was allowed to stand for 1 min. An amount of polyglycol ether blowing agent corresponding to 40 g / t of dry ore was then added, after which the slurry was allowed to stand for a further 1 min. After starting the flotation, a further collector was added stepwise as shown in Table VI below.

The flotation cell was agitated at 900 rpm and air was fed at a rate of 9.0 l / min. Foam-rich samples were collected at time intervals from 0 to 1.0; 1.0 - 3.0; 3.0 - 4.0; 4.0 to 6.0; 6.0 - 7.0; 7.0 - 9.0; 9.0 to 10.0 and 10.0 to 14.0 min after starting the air supply as shown in the table below. Foam concentrate and residual samples were dried, weighed and ground for analysis. They were dissolved in acid and the iron content was determined by direct current plasma spectrometry. From the assay results, yields and concentrations were calculated using standard mass equilibrium formulas. The results are shown in Table VI below in Brochure 35.

33 90 39 8 ΙΛ 3 - m co irv g O ° CNJ - S H m m tn • h w -h ^ ^ ~> -H oJ O O o • h m Π

• H C 4-> o -H

ro -P O (M 10

hoc po m 3 λ: Ό <- a- vo ε 3 r ^ S “° o o w 3 S'- O oo" <jd m cp in o in = r 3 4j n

3 H O, O O

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εΙ ει ει; · ". 71„ 11 .. ·; | ..... ρ 0 | «- mj ro νο | t— C f— f \ l <- g -, τν; · .: So oo. Ω - ρ O nj rvj c \ j 3 po —I * p- ^ ro \ r- ooo. '"ζ r' * w- r — i ^ c. o o o o 9 9 9 θ θ e 1 rr γί rr

Ai m m> 1 Z 2 2 __j im n m 3 o o r — i H JM IM l / l

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o 3 a. o_ a.

O H Q Q Q

:. ·. S.>,? s ~ - - α, υ u u • * * n · i i

2 ö Q Q

Cu aa aa as 34 90 398 en ρ 3

W

C -H

QJ O C \ J t— O -H -H

e -P t "- i— * - 4J -P

'5 h' ä ^ 10 in ti ti? -h a r jV Λ 3 3 M 3 I _ ro rd

• H C O O O CC

+ J 0 0 O

3 λ: 14-1 14-1 M O -H. — I, - (

3. * + J irt o 03 3D

ε e - = r «- eg ui«

3 3 03 in 03 H H

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3 1 (3 * ^ C C

ο 0) O O O G C

3 3 3 μ 3 4-1> 3 3 0 0 0 C 3 3 3 V 3 U1 Ό 4X3 αο, Η Μ -¾ 3 -Η ΜΟ (Μ Ο 4-1 4-4 3 2 5 * 0.. 5 ^

3 3 ^ Ο Ο Ο 5 S

^ 3 C C · Η · Η 05 · 3 · Η C Ό Ό ε I ε I * η EI 0 ίίθ: ο3 ο Γ- | J3-I Η Η .1. Ι 0 Ή γΗ η = r · ro ι en. 3 η η 4-ι ι α> ο | m cH «- εεε £ oi O m cg w χ: .e 3 .. Γ- ^ 3>,>, S o O O 3 .S. $

P »“ H (-H

O> 1> 1 D ^ x

o ^ O 4T 4C CCNOJ

e en O «- O O -hmm

G .V »4« -> «4« NI3CJU

° o o o _ P 03 rfl E Ή f — ι

Ή «H C -H -H

: Q <U 0)

® A a C C C

e -H ®_ ®- C P D

rg rH * -h -p 4J

t? 0 -5 3 UI 4-1 4-1 2 a: z z a: -h -h m it n n oi O o

Λ rH (J O X U U

- in 31 S2 ΰ 3 3 3 3i '3 D O M 3 3 o 2 S S § w x x 5 9, m 2 Γ! M tN ίή S V £ Λ Λ ώ 2 Q- Q- * r * aa aa aa cu ti 35 90398

The results shown in Table VI above show that the monosulfonate collector of the present invention results in a very high yield of good iron in a substantially shorter time than it takes to achieve comparable yields using the disulfonate.

Example 7

Flotation of various oxide minerals

The general procedure of Example 1 was followed, except that various oxide minerals were used in place of the copper ore. All experiments were performed at pH 8.0. The collector used was sodium diphenyloxide monosulfonate disubstituted with branched C12 alkyl groups and 0.024 kg / kg of mineral was used.

36 90398

Table VII

mineral

Mineral yield quartz (SiO 2) 0.204 cassiterite (SnO 2) _0.931 bauxite [Al (OH> 3) 0.989 calcite (CaCO 3) 0.957 chromite i ^ β ^ Γ2θ4) 1,000 dolomite [CaMg (CO 3) 2l 0.968 malachite [Cu 2 CO 3] 2i_0.989_ chrysococcus (CU2H2S i 2Ο5 (OH) 4] 0.6 16 hematite (F6203) 0.97 1 corundum U2O3) 1,000 _rutile (TiO2) _0.970_ apatite [Ca5 (Cl1F) [POt4l3i_0.990_ nickel oxide (NiO ) 0.778 lead (PbS) 0.990 chalcopyrite (CuFeS2l 0.991 chalcosite (Cu2S) 0.993 pyrite (FeS2) 1,000 zinc scintillation (ZnS) 1,000 pentlandite (Nl (FeS)] ® 0.980 elemental-Cu®uu 0.9u AgO 0.873 l! 37 9 0 398

Table VII (continued)

mineral

Mineral intake of barite (BaSOty) _0.968_ molybdenite (MoS2) 0.968 serusite (PbCO3). 0.939 calcite-.i (CaCO 2) Q, 8U7 beryl (BegAlgSiftO ia) _ 0.937 covellite (CuS) 0.788 zirconium (ZrSiOn) 0.876 gratite (C) 0.937 topaz iAl2S10 || (F- | 0H) sc 1 0.95 ^ (CaWOty) _0.871 anatase (T102) o, 909 boehmite (γΑ10.OH) 0.886 diasprore (αΑ10 · 0Η) 0.905 goethite (HFe02) 0.959 20 1. The sample contains some pyrrolite 2. The sample comprises a similar particle size to other mineral samples powdered elemental metal The results shown in Table VII show that: a variety of minerals can be foamed with the collector and method of this invention.

Example 8

Flotation of Molybdenum-Containing Copper Sulfide Alloy Ore - * A series of 30 g samples were prepared from Arizona ore having a particle size of -10 US mesh and:: containing a mixture of various copper oxide and copper sulfide minerals and small amounts of molybdenum minerals.

38 90398

The ore had a copper content of 0.013 and a molybdenum content of 0.000016.

Each ore sample was ground in a laboratory-sized spare mill for 10 s and the resulting fines were transferred to a 300 ml flotation cell.

Each experiment was performed for Malmi at a natural pH of 5.6 in the slurry. The collector was added 0.050 kg / t dry ore and the slurry was allowed to stand for 1 min. Two concentrates were recovered conventionally by hand at intervals of 10 0-2 min and 2-6 min. Just before the start of flotation, a flotation agent, polyglycol ether sold by The Dow Chemical Company under the name Dowfroth® 250, 0.030 kg / t dry ore was added.

In all experiments, the flotation cell was agitated at 15 rpm and air was fed to the cell at 2.7 l / min. The concentrate and residue samples were then dried and analyzed as described in the previous examples. The results obtained are shown in Table VIII below.

39 90398 tn 3 · - in <v)

3 - O O

_ Ui O O O

I I o o o

n n O O O

O O. O

-H O O O

ai -h oj σ> · - + j - = r m ^ r II c o O o 0 3 "c Sl S O o o

•B

E tn

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a in oo «- oo I I H O« - · - 3 0 i. ·> · · 0 -u o o o

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Ti -h s I (0 ^ ^ »n g e tn o o o

H (N

tn 1 13 on co oo: ·. o 3 m 'f »3 0 r r.,: o -m o o o:. , "-J o t— m I -g <M = T ro -: ö § °° =» ^ * n tn O O o '· - - en 3 {n o o in

o \ tn tn (M

e en O O O

C ^ r »e * rf-N

<^ o o o ...: Θ_ Θ

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··· - o 'O O O

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Table Vili (continued)

Cumulative metal intake 5 and content

Dosage Cu- Cu-pl- Mo-Mo-pi-Collector (kg / t) intake other intake L, D-CioDPO (SO3Na) 1 © 0.0 5 0 0.90 5 0.16 1 0.917 0.000040 10 L> D-CioDPO (SQ3Na) 2® 0.050 0.598 0, 129 0.745 0.000024 L, D-CioDPO (SO3Na) 1® 0.025 0.765 0, 1 * 43 0.812 <\ 000025 1. Disubstituted with branched C12 alkyl groups Sodium diphenyloxide monosulphonate 15 2. Sodium diphenyloxide disulphonate disubstituted with branched C12 alkyl groups

The results shown in Table VIII show that the monosulfonated collector 20 of this invention provides significantly improved yields of better copper and molybdenum than the corresponding disulfonated collector.

Example 9

Ha11imond Tube Foaming 25 The procedure described in Example 1 was followed using a number of different mineral nests and different collectors. Foaming concentrates and residues are subjected to metal determinations using acid leaching and direct current plasma spectrometry. The results are shown in Table IX below. Although the results are presented in one table, it is important to note that the results for each mineral were obtained separately. In each case, the foaming was carried out at the inherent pH of the ore in the respective slurry form, i.e. at a pH of 5.8 for rutile-35 Ia; 6.7 for apatite; 6.0 for pyrolusite and 6.8 for the diaspore.

li 41 90398

Table IX

Apa- Pyrolu- Dia-Dosage- Rutile-titite-spore - Collector (kg / t) intake intake intake intake 1 B, D- 0.0001 0., 021 0.009 ~ C, 2DP0 (SO3Na), ® 2 BD- 0.0005 0.323 0.038 C, 2DP0 (SO3Na), ® 3 B, D- 0.0010 0 ^ 713 0.463 C | 2DP0 (SO3Na) t® 4 BD- 0.0100 0.954 0.856 0.745 0.598 C, 2DP0 (SO3Na), ® 5® B, D- 0.0001 0.000 0.000 C, 2DP0 (SO3Na) 2® 6® B, 0- 0.0005 0.015 0.007 C, 2DP0 (SO3Na) 2® 7® B, D- 0.0010 0.087 0.297 C, 2DP0 (SO3Na) 2® 8® BD- 0.0100 0.175 0.518 0.314 0.280 C12DP0 (SO3Na) 2® 9® BD- 0.0500 0.371 C, 2DP0 (SO3Na) 2® 10® BD- 0.1000 0.815 0.849 C, 2DP0 (SO3Na) 2® 42 90 39 8

Table XX (continued)

Apa- Pyrolu- Dia-

Dosage Rutile rite sperm sporangia

Experiment Collector (kg / t) intake intake intake intake 11 BM- 0.0001 0.000 0.000 C, 2DPO (SO3Na), ® 12 8, M- 0.0005 0.011 0.000 C, 2DP0 (SO3Na), ® 13 B, M- 0.0010 0.034 0.111 C12DPO (SO3Na), ® 14 B, M - 0.0100 0.129 0.277 0.289 0.166 C12DPO (SO3Na), ® 7 15 BM- 0.0500 0.296

Ci2DP0 (SO3Na), ® 16 BM- 0.1000 0.644 0.680 C, 2DP0 (SO3Na), ® '17® BM- 0., 0001 0.000 0.000 C, 2DP0 (SO3Na) 2® 18® BM- 0.0005 0.000 0.000 C, 2DPO (SO3Na) 2® 19® BM- 0.0010 0.000 0.000 C, 2DP0 (SO3Na) 2® '7 20® BM- 0.0100 0.009 0.011 0.017 0.005 C12DP0 (SO3Na) 2®' 21® BM- 0 .0500 0.027

C12DP0 (SO3Na) 2® J

22® B.M- 0.1000 0.065 0.081 C, 2DP0 (SO3Na) 2® 'li 43 90398

Table IX (continued)

Apa- Pyrolu- Dia-

Dosage Rutile titite spore spore- K ° e Collector (kg / t) intake intake intake intake 23 L, D- 0.0001 0.104 C10DPO (SOjNa) i® 24 L, D- 0.0003 0.310

CioDPO (SOjNa), ® 25 L, 0 - 0.0005 0.563 C, 0DPO (SO3Na), ® 26 L, D- 0.0010 0.869 C10DPO (SO3Na), ® 27 L, D- 0.0100 - 0.773 0.605 C10DPO (SO3Na), ® 28 LD- 0.0200 - 0.956 CI0DPO (SO3Na), ® '29® L, D- 0.0001 0.030 C10DPO (SO3Na) 2® 30® LD- 0.0003 0.041' C10DPO (SO3Na) 2 ® 31® LD- 0.0005 0.095 C, 0DPO (5O3Na) 2® '32® LD- 0.0010 0.164 C, 0DPO (SO3Na) 2® 33® LD- 0.0100 - 0.444 0.248 C | 0DPO (SO3Na) 2® 34® LD- 0.0200 - 0.581 C, 0DPO (SO3Na) 2® 35 LM- 0.0005 0.051 C, 0DPO (SO3Na), ® 36 LM- 0.0010 0.120 C) 0DPO (SO3Na), ® 44 903S8

Table IX (continued)

Apa- Pyrolu- Dia-Dosage Rutile tite rite Spori K ££ Collector (kg / h) intake intake intake 37 lM- 0.0015 0.559 C, 0DPO (SO3Na), ® 38 L, M- 0.01 - 0.235 0.267 C10DPO (SO3Na), ® • 39® L, M- 0.0005 0.01 1 C, 0DPO (SO3Na) 2® 40® L, M- 0.0010 0.21 C, 0DPO (5O3Na) 2 © 41® L, M- 0.0015 0.041 C, 0DPO (SO3Na) 2® 42® LM- 0.0100 - 0.005 0.005 C10DPO (SO3Na) 2® 43 L, D- 0.0100 0.744 - 0.889 C, 6DPO ( SO 3 Na), 10 44® LD- 0.0100 0.289 - 0.522

C 6DP0 (S03Na) 2 H

45 L.M- 0.0100 0.185 - 0.348 C16DP0 (SO3Na), 12 46® L.M- 0.0100 0.109 - 0.176 _C16OPO (SQ3Na) 213 * _ 47 L.D- 0.0100 - - 0.733

Ci2DP0 (503Na) i1 (1 48® L.D- 0.0100 - - 0.337 C, 2DPO (SO3Na) 215 11 45 90 39 8

Table IX (cont'd) 1. Sodium diphenyloxide monosulfonate disubstituted with branched C12 alkyl groups 2. Non-embodiment of the invention C 3. Sodium diphenyl oxide disubstituted with branched C12 alkyl groups 10-substituted natriumdifenyylioksididisulfonaatti 6. linear di-substituted with C10 alkyl natriumdifenyylioksidimonosulfonaatti 7. linear di-substituted with C10 alkyl natriumdifenyylioksididisulfonaatti 15 8. The linear C10 alkyl group mono-substituted with natriumdifenyylioksidimonosulfonaatti 9. The linear C10 alkyl groups monosubstituoi natriumdifenyylioksididisulfonaatti TU-10 linear C16-disubstituted with alkyl groups of 11 to 20 natriumdifenyylioksidimonosulfonaatti Sodium diphenyloxide disulphonate disubstituted with linear C16 alkyl groups ti 12. Sodium diphenyloxide monosulfonate monosubstituted with a linear C16 alkyl group 25. Sodium diphenyl oxide disulfonate monosubstituted with a linear C16 alkyl group 14. Sodium diphenyl disodium disubstituted with linear C12 alkyl groups

The results shown in Table IX above show that the monosulfonated collector used in the process of this invention regularly provides higher yields of different minerals compared to collectors that are similar except for monosulfonation.

Example 10

Sequential flotation 5 In this example, that described in Example 1 is used

Hallimond-putkivaahdotusmenettelyä. In each case, the material to be fed was a mixture of the components listed in Table X below in a weight ratio of 50:50. The collectors used in each case and the mineral-10 additions obtained are also shown in Table X below.

47 90 3 98

CN

* -p I h oq cr \ cm o t-m t— m m ^ o O't— t — cn

C O -P M3 OO OO C— CT »fO v £ J O CM MD CM t— CO CM

5 & £ o = r r-vo o co comd cm co cm co r- cm ωο cj * 'i' * 'c - f Γ i' f (t Γ

p ^ c QO OO OO OO OO OO OO

r-H

3 3

P rH

s * • h I -h ^ rt ~ - ud co ro = r = rcr> roc— <- »—t— S op * —ϊτ cm σ> t— oco cr> zr o · - oo- = rg * p voctn t— σ> aa cr \ md oo oo cr> co cr \ oocr> yj 2 OO oo oo oo oo oo oo (N Ή · γΗ -Η · Η Ή Ή · Η -Μ ^ 4 4- * 44 4- * +4 Ή Ή 4-4 4-4 · Η * γΗ 4-4 4-4 · Η Ή +4 4- »4-4 4-4 4-Ι-Ρ 4-4 4-4 4> 44 - Ρ4- * 4> 4> • Η · Ρ · Η Ή + J +4 · Η -Η 44 44 · Η Ή 44 4-4 I · Η Ή -Η · Η 'Μ · Η · Η Ή -Η Ή · γ4 · Η -Η · Η -Η ο 4-4 -Ρ 44 £ £ Ή · Η WU) -Η -Ρ W 1Λ · Η · Η α 4> ro rd ΟΟ 4> 4-> 4 * 4 * ί 4 4 ^ Λί £ £ £ C £ £ * Η rl ΡΡ 3 3 ΡΡ 3 3 -C, 3 ο α> oo oo ora 33 3 3 3 3 ο ο c χ: χ: T3TJ eg χ χ; ee χι χ χ χ ο ο

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M OO C4C4 a P 0 3 -H -H

3 X! C 33 33 3-3 33 PP PP Cn X

3: * .. e- ·

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- 2 ^ tn o m o eno LO o ur »o in o eno

U) x CM O CM O CM O CM O CMO CM O CM O

O \ o · - O «- O« - O r ~ O'- o · - o - C CT1 · ** · "* r * ri 4 ~» f> r * '* »·, / \ ^ ^ 00 00 00 00 00 00 00 0Θθθβθβθθ9 (^ © θ0 n) n) (On} (Oci id n) id id id idllni 2S3S SS SS ZZ SS SS ZZ SS SS zz JP rn mrr »m (n mm mm mm mm 00 00 00 00 00 00 00 10 co co cn co to co co to co coco coco 00 00 00 00 00 00 00 rO 0-0- 0-0- 0-0- 0-0- 0-0- 0-0- 0-0- r ~ Ω q qq aq qq aa aa ao 0 OO OO Γ4 t> 4 (N n <n <nm «nm rs # _2 υο o cj oo ou ό <S ocj oo

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48 90398 Ή ρ • H Ρ Ρ Π3 P to

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c o 4J co ao t— as r— o> vooo II ζ • h 1¾ .μ ^ r *** · - f ^ "T? secoooooooo C 3 S ϊ c« 2

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k "1 C I · Η τ-ί · Η G ^ tO (0 r-f rH T3 jrrt

• H 0 4-> -H -H H H 03 ro -H -H

S 4-) 10 (0 OO Π3 03 -H -H: rcj h £ cx ^ xiM ^ a (al4J-p ri -h OCD-HH> ί> i OO DG i — ic ^ c en t? 0 (0 ^^ -) 4-1 μ μ -n £ e> 1

rG U

Ui ‘-p4 d« mo loö rno 1X10 h 7.

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II

49 90398

The results shown in the table above show that the various minerals to be foamed by the method of this invention can be effectively separated by controlling the accumulator dosage. For example, it is clear that although both apatite and dolomite can be foamed by the method of this invention, apatite rises more easily to the surface at lower aggregate doses than dolomite. Thus, apatite can be foamed in the first stage, where the dosage is low. This may be followed by a second flotation at higher collector doses to flotate the dolomite. As a review of other experiments in this example shows, similar distinctions can be made for other minerals.

Example 11 15 Separation of apatite from quartz and dolomite

The procedure was otherwise as in Example 4, but the samples contained 30% apatite, 60% quartz and 10% dolomite. In Experiments 2 and 3, further purified hydrocarbon was added. The results obtained are shown in the following tau-20 lock XI.

50 90398

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H 4-> Ό ΓΗ • 3 -3 o t o oj m ui w S C * <M O VO £

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J-1 CM t— C-- OO DU) c vo (\ i - t ~ - * j a c% ^ ® ^ o 3 § ΐ o O O O Λ-hc

Cm 3 O 0) U) D. ϋ • H U) O TO> 1 3

U) 4-1> 4J

S — OOOOOOO 3 ίπ E'§ u) 4J m m m in »- m» - e -1 3 .c ° tr o 00 00 00 H? £ tl 2 ^ ^ ^ r '^ £ O Π3 <c5o 00 oo 00 _t | · §, π> M M -.3 O -3 cd u:

3 Ή 4-1 M

ui> 4 o a (n) CÄ>, 3 3

^ C Ai 4-> AI

- ^ © CD -—I -H

- ra -H G 3 cc Z <0 © Z® c 3 1-33 J? Z J? -3 3 O -3 -rl o m-3 0-3 -3 AI r3 U) o «Λ O 3 ΙΛ C OU) 3 O 00 O νΛ 3 λ-3 ft OGM Ai y (3 hi" 3 ft M 3 30 oiy »30 cm>;. * N QJOQ 0 3 Λ!: 0, —1 OOQm cmvj>, c -3 3-3

Oh. O 11) n 11 n 3 C U) 4-> 4->

Ai Y. 7 ϋ 'V Ä · h D -3 -3 4J U)

O Q J Q 04-1 U) M CD CD

^ ί Q 3> 3 3 M 3 4-> u) -J. 4J 00 4J '4-1 CD 3 UI -3 —1 4- * 4-1 U) ^ (D03 3 CD OCM Ai 4-1 4-1 M -h -h 3 -h U) u) 3 MJ 3 ) 3 o O 13 c -—f * —1 · · · ro α> Γ CM 3 cn ^ ® ~ ”m% £ lc tflj: iX3

P

II

si 90398

The results shown in the table above demonstrate the ability of the collector of this invention to foam apatite prior to dolomite, i.e., to separate apatite and dolomite. Experiment 4, which corresponds to the industry standard procedure, does not result in a comparable separation of apatite and dolomite, and thus results in the recovery of phosphorus significantly contaminated with magnesium. The addition of a hydrocarbon to the process according to the invention results in a higher concentration of phosphorus at a slightly lower yield, while the amount of magnesium recovered decreases.

Example 12

Apatite flotation

The procedure was otherwise as in Example 11, but the foamed ore was a mixture containing 30% apatite, 10% calcite and 60% quartz. The results obtained are shown in Table XII below.

52 90 3 9 8

Table XII

^ Dosage

Test Collector (kg / h) p-intake p-content 1 LD-ClODPO (SO3Na), ® 0.050 0.317_0-> 128 2 L, DC, oDPO (SO, Na), ® 0.100 0.792 0.137_ 3® oleic acid 0.100 0.551 0.064 10 1. Sodium diphenyloxide mononosulfonate disubstituted with linear C10 alkyl groups 2. Non-Invention Embodiment 15 The results shown in Table XII demonstrate the efficacy of the method of the invention in apatite recovery. A comparison with Example 11 also shows that the dosage required to achieve a given intake depends on the minerals to be foamed in each case.

20 Example 13

Flotation of carbon-based inks Five sludges were prepared by forming a pulp containing in each case 240 g of printed paper (70% newspaper and 30% magazine); 1.61 g of diethylenetriaminepentaacetic acid (color adjusting agent); 10.65 g sodium silicate; the number of collectors shown in Table XIII; and 5.64 g of hydrogen peroxide, using a sufficient amount of water to obtain a slurry having a solids content of 2% by weight. The pH of the slurry was 10.5, unless otherwise stated, and the temperature was 45 ° C. The preparation of the pulp took 30 min. Each slurry was prepared from exactly similar sides to ensure that the amount of ink was comparable in each of the five slurries prepared.

Il 53 90398

The pulp-formed slurry was transferred to a 15 L Voith Flotation Cell, diluting with sufficient water to fill the cell. An amount of calcium chloride was added to the pulp to a water hardness of 5,180 ppm CaCO 3. Flotation was started by feeding an air bubble through a vigorously stirred mass and continued for 10 min. The foam was then removed by conventional manual means to give a flotation product.

The flotation product was then filtered and dried. The flotation product was analyzed colorimetrically using a stepwise composition scale of 0-10, where 0 is completely white and 10 is completely black. Pulp mats made from the contents of the cells were examined with a powerful microscope to determine the ink particles per unit area.

The results obtained are shown in Table XIII below. Conditions were similar in each experiment unless otherwise noted.

54 90398 \

•B

4-) I -U> 3 3 CO 3 03 3 Ο! Γ-l 4-> β O <D CO 0

-H '4-) 0 3 3 3 <4H

3 CO. 4-> M <#> M <#> M OP <- ».-H} 0 0 3 CO CO CO 3

rH. 4-)> · -H 3 3 m it O ram CO

aj) (T3M CD £ H <N «H m« nr- 0

U) '£ 3 G

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ft ΙΛ H π3 ti ra ra ra C 3 3 0) 03 -H (0 a) 3raa) <a eee-nge m 4J x3^ 4-> 3 d HE ^ 4 £ H> £ Ö ro CO ^ 03 \ 10 CO 3 3> h 3 3 3> i 3 3 33 3 -H ·· \ 3 -HW 3 3 Ä + JÄ Ä 4-> Ä> g * T3 1 — f »- ·

\\ φ O -H 3> 1 (J

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X 3 W © © 0) 333 o -S S 0 m o .g .ti u λ; λ y _ „/ ^ 4J c« o

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3 · ». -r-t f — I

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M is> o o ^ o e -H & c cm t \ i cm cm '3 Jj 3 c

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<3 Ή 4-> CO

ω>. -π 3 _ _ _> »r — I 4-)

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0 ^ 3 3 3 hi 3 G4 3 * ·. 3033

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0 ^ 2 Q Q Q: O ^ SS

3 ° | 2 2 2 £ 'S ϋί ϋ

O 2 U UVJ-rH-HDO

* V ά ώ ά J2 £ S * Ο. _ / _ί _ί 4) 3 Τ3 s3 _ι ^ <υ χ: w a 3 co • Η · Η 3? 3 ω j> η Q ^ ^ H M ro <υ I Φ Θ Θ ο cm pri in xl li 55 90398

The results shown in the table above show that the method of the present invention is effective in separating graphite ink and other carbon-based inks from paper by deinking the recovered paper by foaming. A comparison of Experiments 2-5 to Experiment 1, which is close to the standard procedure currently used in the industry, shows that the use of the collectors of this invention results in a higher ink yield with a significantly lower collector dose.

Claims (18)

56 90398 A process for recovering minerals by foaming, characterized in that the aqueous slurry containing particulate minerals is foamed in the presence of a collector comprising alkylated diaryloxydisulfonic acid or a salt thereof or mixtures of such acids or salts with at least 20% of sulfonic acid or their salts. monosulfonated, under conditions such that the minerals to be recovered foam. Process according to Claim 1, characterized in that the monosulphonic acid or its salt corresponds to the formula (RU (SO, -M +) y (SO 3 -M +) x 20 in which each group R is an independently saturated or unsaturated alkyl group or a substituted alkyl group; m and n are independently 0, 1 or 2, each group M is independently hydrogen, an alkali metal, an alkaline earth metal or an ammonium or substituted ammonium group, and each of x and y is independently 0 or 1, provided that the sum of x + y is 1. Process according to Claim 2, characterized in that R is a 3-alkyl group having 1 to 24 carbon atoms. Process according to Claim 2 or 3, characterized in that R is an alkyl group having 6 to 24 carbon atoms. . Process according to Claim 2 or 3, characterized in that R is an alkyl group having 10 to about 16 carbon atoms. Il 57 9 0 3 9 8 Process according to Claim 2 or 3, characterized in that R is a straight-chain or branched alkyl group. Method according to one of Claims 1 to 6, characterized in that the sum of the numbers m and n is 2. Method according to one of the preceding claims, characterized in that the total concentration of the collector is at least 0.001 kg / t and at most 5.0 kg / t. Process according to one of the preceding claims, characterized in that at least 25% of the sulfonic acid or its salt is monosulfonated. Process according to one of Claims 1 to 8, characterized in that at least about 40% of the sulfonic acid or its salt is monosulfonated. Process according to one of Claims 1 to 8, characterized in that at least about 50% of the sulfonic acid or its salt is monosulfonated. Method according to one of the preceding claims, characterized in that the recovered mineral comprises graphite and the aqueous slurry comprises a pulp made of paper. 13. A collector composition useful in the recovery of minerals by foaming, characterized in that it comprises alkylated diaryloxydisulfonic acid or a salt thereof or mixtures of such acids or salts in which at least 20% of the sulfonic acid or its salts have monosulfonic acids or their salts of formula 30 0-) m (h) „0-0 35. P (SOjMMy (SOjM * -) * 58 9 0 3 9 8 wherein each group R is independently a saturated or unsaturated alkyl group or a substituted alkyl group; each of m and n is independently 0, 1 or 2; each group M is independently hydrogen, an alkali metal, an alkaline earth metal or an ammonium or substituted ammonium group, and each of x and y is independently 0 or 1, provided that the sum of x + y is 1. Composition according to Claim 13, characterized in that R is an alkyl group having 1 to 24 carbon atoms. Composition according to Claim 14, characterized in that R is an alkyl group having 10 to 16 carbon atoms. Composition according to Claim 13 or 14, characterized in that R is a straight-chain or branched alkyl group. Composition according to one of Claims 13 to 16, characterized in that the sum of the numbers m and n is 2. Composition according to one of Claims 13 to 17, characterized in that the total concentration of the collector is at least 0.001 kg / t and at most 5.0 kg / t. li 59 90 398