Flotation and depression control of arsenopyrite through pH and pulp redox potential using xanthate as the collector

The role of pH and pulp redox potential (E_H) to control the flotation and depression of arsenopyrite has been investigated through studies on microflotation of arsenopyrite crystals and batch flotation of an arsenopyritic ore using isopropyl xanthate as collector. The transition between flotation and depression of arsenopyrite is established by the reversible potential of the xanthate/dixanthogen couple. Adsorption of arsenate ions on ferric hydroxide has been studied through electrokinetics to delineate mechanisms involved in the depression of arsenopyrite using oxidants. Chemical binding between arsenate species and ferric hydroxide sites on arsenopyrite is suggested as the mechanism responsible for depression of arsenopyrite. E_H conditions are given for the flotation and depression of arsenopyite at various pH values for the arsenopyritic ore.     

Flotation of germanium n and p with potassium ethyl xanthate

Steady-state potential of the germanium electrode, flotation tests and results of spectrophotometric ATR (Attenuated Total Reflection) measurements of germanium surface in the presence and absence of potassium ethyl xanthate are presented.On the basis of the steady-state potentials of the germanium electrode in ethyl xanthate solutions, the standard potential E⁰ of the reaction: Ge + 2EtX^? = Ge (EtX)₂ + 2e is estimated. The pH ranges of the dixanthogen (EtX)₂ and the germanium xanthate Ge(EtX)₂ species predominance in the bulk solution are calculated. It has been established that flotation of germanium is greater than the natural floatability (in the absence of collectors) in the pH ranges where (EtX)₂ or Ge (EtX)₂ are present in the bulk solution. Spectrophotometric results reveal the presence of (EtX)₂ and Ge (EtX)₂ on the surface in the same pH ranges as is calculated for the bulk reactions.No significant differences in the surface properties and the flotation behaviour between germanium n and p have been found.     

Formation and recognition of alkyl xanthyl thiosulphates in sulphide ore flotation liquors

Alkyl xanthyl thiosulphates (R.OCSS.S₂O₃^?) (RXT^?) are formed in solution by mild oxidation (e.g. by I₂) of solutions containing both xanthate and thiosulphate. They can also be formed by reaction of Cu²⁺ with xanthate and thiosulphate, reaction of dixanthogen with thiosulphate, and by reaction of xanthate with tetrathionate; these last three reactions can occur in flotation pulps in slightly acid or alkaline solutions (pH 5–10).Alkyl xanthyl thiosulphates are stable in acid and neutral solution; the solutions have a UV absorption maximum at 289 nm. In strongly alkaline solution (pH 12) RXT^? decomposes within a few minutes to yield a xanthate (mostly) plus a little perxanthate. At pH 10 this decomposition to xanthate takes about 48 h. At pH 7–9 RXT^? is relatively stable. RXT^? is not extracted from aqueous solution with common solvents (chloroform, iso-octane, cyclohexane, or ether). It forms a water-insoluble adduct with cetyltrimethyl-ammonium bromide (CTAB); this adduct can be extracted into chloroform, and the extract has a UV absorption maximum at 296 nm.RXT^? was found in solutions from the gangue-sulphide flotation section at Renison Ltd, the zinc flotation circuit and the copper flotation circuit at Mount Isa Mines Ltd, and the lead flotation section of The Zinc Corporation Ltd. The presence of RXT^? in operating flotation plants has various practical and theoretical implications.     

Studies on the flotation of sulphides. II. The effect of urea on the flotation of zinc sulphide

The effect of three hydrophilic non-electrolytes on the flotation of ZnS has been studied using a Fuerstenau microflotation cell. The rate of flotation increases with increasing concentrations of urea(OC(NH₂)₂), but is unaffected by thiourea and slightly decreased by sucrose. This effect of urea was observed with all samples of ZnS (synthetic and natural) studied, is unaffected by extensive purification of the urea and by the presence or absence of O₂, and is reversed by dilution. Urea also increases the rate of flotation (in a Partridge and Smith cell) in the absence of frother. Surface tension measurements showed the absence of detectable concentrations of surfactants in the urea. The presence of urea has no effect on the rate of sedimentation (i.e. on the degree of aggregation) of ZnS or on the volume of water carried over during flotation. It is suggested that the effect of urea is (at least in part) to increase the rate of bubble-particle attachment by modifying the structure of water.     

Adsorption of barium ions by β-manganese dioxide and its activation in oleate flotation

The Ba²⁺ ion adsorption isotherms on β-MnO₂ were of the Langmuir type. The endothermic heat of adsorption (40 kJ mol^?1) is ascribed to entropy contributions associated with the $\text{Na}{}^{\mathrm{+}}\text{Ba}{}^{\mathrm{2+}}$ ion-exchange mechanism. The Ba²⁺ ion adsorption density was higher at pH 10 than that at pH 7, due to the more negative surface charge at the higher pH. Ba²⁺ ions were found to reverse the sign of the ζ potential of the MnO₂ particles.More oleate was adsorbed by β-MnO₂ in the presence of Ba²⁺ ions than in their absence. The oleate adsorption isotherms on Ba²⁺-activated MnO₂ were of the Freundlich type and indicated an exothermic process. Hallimond flotation recovery of Ba²⁺-activated MnO₂ was higher at pH 10 than at pH 7, although less oleate was adsorbed at the higher pH. At pH 7, Mn²⁺-activation led to higher recoveries than Ba²⁺-activation. It seems that the attraction between the surface and the activator plays an important rôle in determining the flotation recovery.     

Optimisation of operating variables of fine coal flotation using a combination of modified flotation parameters and statistical techniques

Optimisation of flotation parameters using rate models is not a new concept. The kinetic model based on time recovery data, which uses the extra dimension of rate, has been in vogue since time immemorial for scaling up of lab data. Often, interpretation on the performance of a flotation circuit, based only on R_∞ (the ultimate recovery) and k_s (the first-order rate constant of the component) may lead to wrong conclusions. In such cases, a modified flotation rate constant k_mod defined as the product of R_∞ and k_s, i.e., k_mod=R_∞k_s and selectivity index (SI), defined as the ratio of the modified rate constant of valuables to the modified rate constant of gangue, can be used. An attempt has been made in this paper to optimize the batch laboratory froth flotation parameters of fine coal using the above two concepts, i.e., k_mod and SI and statistical techniques.A flotation bank containing four Outukumpu cells was optimized using the results obtained from the lab study. The airflow number and the froth number were used as a basis for scale up. To gauge the performance of the froth flotation circuit, an efficiency parameter called the coefficient of separation or the CS was used. The yield from the flotation circuit improved, the froth ash reduced and the rejects ash went up.     

The collectorless flotation of chalcopyrite ores using sodium sulfide

The collectorless flotation process has been tested on six different chalcopyrite ores while monitoring the potentials ($\text{E}{}_{\mathrm{<mtext>h</mtext>}}$) of the pulp. The results show that collectorless flotation is effective only under oxidizing conditions. In addition, the flotation requires that the chalcopyrite surface be relatively free of hydrophilic oxidation products, which can be accomplished by treating the ore pulp with sodium sulfide (Na₂S). On the basis of these findings, methods of improving the collectorless flotation process are discussed.     

Interactive effects of the type of milling media and CuSO4 addition on the flotation performance of sulphide minerals from Merensky ore Part II: Froth stability

The limitations of pulp chemistry measurements in the flotation of a platinum group mineral (PGM) bearing Merensky ore were demonstrated in Part 1 of this article. In this paper the importance of the contribution of the froth structure due to changing froth stability is analysed using the batch flotation data. The effects of mild steel (MS) and stainless steel (SS) milling media and the addition of copper sulphate on the flotation performance of the sulphide minerals in Merensky ore have been evaluated in relation to the changes in stability of the froth phase. The effects of pulp chemistry and froth stability on the flotation of sulphide minerals were distinguished by using two different rate constants (k_t and k_w). The rate constant (k_w) calculated as a function of cumulative water recovery was used to describe characteristics of froth phase and k_t was calculated as a function of flotation time. The results revealed that the type of grinding media and copper sulphate addition had an interactive effect on the froth stability. While mild steel (MS) milling increased the froth stability due to the presence of hydrophilic iron hydroxides and colloidal metallic iron, the addition of copper sulphate reduced the stability, especially with stainless steel (SS) milling. Copper sulphate addition had a dual role in the flotation of Merensky ore in that it caused destabilisation of the froth zone as well as activation of selected sulphide minerals. The dominant effect was found to depend on the type of milling media and floatability of the mineral in question and this work has demonstrated the importance of using a combination of measurements to evaluate flotation performance holistically.     

Flotation of calcareous Mussorie phosphate ore

Separation of Mussorie rock phosphate (P₂O₅ = 20%) from Uttar Pradesh, India, containing pyrite, calcite and other carbonaceous impurities by flotation has been successfully attempted to upgrade the phosphate values. Based on Hallimond cell flotation results of single and synthetic mineral mixtures of calcite and apatite using oleic acid and potassium phosphate, conditions were obtained for the separation of calcite from apatite which is considered to be the most difficult step in the beneficiation of calcareous phosphates. Further studies using 250 g of the mineral (?60 +150 and ?150 mesh fractions, deslimed) in laboratory size Fagergren subaeration machine employed a stagewise flotation viz. carbonaceous materials using terpineol, pyrite using potassium-ethyl xanthate and calcite using oleic acid respectively. Separation was, however, found to be unsatisfactory in the absence of a depressant.Among starch, hydrofluosilicic acid and dipotassium hydrogen phosphate, which were tried as depressants for apatite in the final flotation stage, dipotassium hydrogen phosphate proved to be superior to others. However, the tests with the above fractions did not yield the required grade. This was possibly due to insufficient liberation of the phosphate mineral from the ore body and different experimental conditions due to scale up operations. Experiments conducted using ?200 mesh deslimed fractions has yielded an acceptable grade of 27.6% P₂O₅ with a recovery of about 60%. The results have been explained in terms of the specific adsorption characteristics of phosphate ions on apatite and the liberation size of the mineral.     

Effects of an ultrasonic field on the flotation selectivity of barite from a barite-fluorite-quartz ore

The influence of an ultrasonic field frequency of 22 kHz and intensity of 0.5 W/cm² has been examined on the flotation of barite, fluorite and quartz as well as on their ability to adsorb dodecyl- and cetylsulphates. The tests affected the flotation of pure minerals in a Hallimond tube as well as the flotation of a natural barite-fluorite ore. It has been found that ultrasonic pretreatment of minerals causes an increase in the flotation rate of barite and a decrease in the flotation rate of fluorite. Ultrasonic pretreatment of the natural barite-fluorite ore before its flotation is the most advantageous. In this case it is possible to obtain concentrates of barite composed of a low amount of CaF₂. This may be explained by the different effects of ultrasonic vibrations on barite and fluorite which cause some improvement of the selectivity of their flotation separation.For a full explanation of these results the effects of ultrasonic pretreatment of the minerals on their adsorption properties and surface topography have been investigated, as well as the direct influence of ultrasound on the adsorption process.     

Selective flotation of scheelite from calcium minerals with sodium oleate as a collector and phosphates as modifiers. I. Selective flotation of scheelite

The effects of Na₂SiO₃, Na₃PO₄, Na₄P₂O₇, (NaPO₃)₆, quebracho, tannic acid and S 808 (sulphonated product of rough phenantrene) on the floatability of the following five pure minerals: scheelite, calcite, fluorite, garnet and quartz, with sodium oleate as collector were investigated in detail as well as the role of pH on these effects. The results obtained indicate that Na₄P₂O₇ and (NaPO₃)₆ were effective modifiers for the selective flotation of scheelite. The results of the batch flotation tests on mixtures of these minerals showed that the recovery of scheelite from scheelite-silicate mixtures (31% WO₃) with (NaPO₃)₆ or Na₄P₂O₇ increased by 20% as compared with sodium silicate and the WO₃ grade of the concentrate by 5%. At room temperature, the scheelite-calcium mineral mixtures could not be separated with sodium silicate. In the separation of these mixtures with the phosphate modifiers, a concentrate grade of 47–60% WO₃ was obtained at 70–90% recovery. This showed that the flowsheet of the selective flotation of scheelite with phosphate modifiers may replace the conventional Petrov's process.     

Perxanthates — A new factor in the theory and practice of flotation

A compound with a UV absorption maximum at 348 nm was observed in Mount Isa copper flotation plant solution. This spectrum was similar to that of the product of reaction of xanthate and peroxide in dilute, alkaline aqueous solution. The compound was termed perxanthate (more correctly “O”-alkyl dithiomonoperoxycarbonate).A new compound, ammonium sec-butyl perxanthate (C₄H₉ OCSSO·NH₄), was prepared by reacting potassium sec-butyl xanthate and hydrogen peroxide in dilute alkaline solution, acidifying, extracting into iso-octane, and precipitating with anhydrous ammonia. Solutions of this compound were compared with solutions containing the Mount Isa compound. Each compound was found to have the same UV absorption spectrum in a given solvent (alkaline aqueous, acid aqueous, chloroform, iso-octane, iso-amyl alcohol, and n-butyl acetate), but the spectra were different in different solvents (especially in alkaline and acid aqueous solutions). Both compounds could be extracted from acid, but not alkaline, aqueous solutions by organic solvents, and both had similar IR and mass spectra.It was concluded that the perxanthate in plant solution resulted from reaction of xanthate with peroxide derived from reduction of oxygen during flotation. This lends credence to the electrochemical theory of flotation and has some important theoretical and practical implications.     

Efficient flotation recovery of lead and zinc from refractory lead-zinc ores under low alkaline conditions

In this study, a new flotation approach, a low-alkaline and non-desliming process, was introduced for improving lead and zinc recoveries, lowering production cost and reducing environmental pollution. Lab-scale experiments results show that the new process contributed to the flotation of the complex mixed sulfide-oxide lead and zinc ore regarding two aspects: (1) High alkaline process (pH = 12±) was replaced by low alkaline process (pH = 9±) by using collector WS (a mixture of ethyl thiocarbamate, ammonium dibutyldithiophosphate and dithiophosphate-25) and combined depressant Na₂S/ZnSO₄/Na₂SO₃ for lead sulfide flotation; (2) Non-desliming process was successfully achieved by using collector MA (a mixture of ether amine, hydroxyethyl cellulose and polyacrylic acid) and combined depressant SHP/SS (sodium hexametaphosphate/sodium silicate) for zinc oxide flotation. And 43.37% Pb in the Pb concentrate was recovered, the corresponding Pb grade was 52.73%, total 84.42% Zn was recovered by the flotation of zinc sulfide minerals and zinc oxide minerals. Moreover, the two aspects of the new approach were systematically verified from lab-scale to industrial-scale application. The industrial-scale flotation tests show that Pb recovery in Pb concentrate increased by 1.86% compared with that of original system during industrial-scale tests period, and the Pb recovery increased by 4.09% compared with that of original system before industrial-scale tests period, while the Zn operating recovery in zinc oxide concentrate improved by 19.52%. Moreover, the total reagent cost of the whole new process significantly declined by 3.93 yuan per ton of ore.     

Flotation study on high-hercynite ilmenite ores

Ilmenite from Polish magnetite-ilmenite ores has been floated with tall oil. Chemical and mineralogical analysis of the flotation products established that the ores containing hercynite are difficult to upgrade. For ore A (18.7% hercynite and 32.8% ilmenite) and ore B (9% hercynite and 41.3% ilmenite) poor concentrates containing much less than the required 45% TiO₂ have been obtained. It has been found that the higher concentration of hercynite in the feed, the lower the grade of concentrates in respect to TiO₂ is obtained. Ilmenite ore containing a minor amount of hercynite (ore C, 0.15% hercynite and 12.3 ilmenite) gives a good concentrate. By means of microscopic observation it was established that hercynite floats together with ilmenite in the rougher flotation but is partially depressed in the cleaning and scavenging flotation.     

Mineral processing and extraction of rare earth elements from the Wadi Khamal Nelsonite Ore, Northwestern Saudi Arabia

A technological sample (50 kg) from Wadi Khamal Nelsonite ore was subjected to magnetic and flotation concentration techniques. Excellent recovery percentages of 72.95% and 71.22% were achieved by the dry/wet magnetic and flotation concentration techniques, respectively. The weight of the apatite concentrate reached a reasonable percentage of approximately 23.5% with an overall 40.23% P₂O₅ total content. Analytical data of the apatite concentrate after digestion in concentrated sulfuric acid revealed that the total content of the rare earth elements (REE) constitutes about 0.2% of the total apatite content. The REE content (0.2%) was partitioned between phosphoric acid liquor (65%) and gypsum precipitate (36%). The extraction of the REEs from the phosphoric acid liquor using oxalic acid and sodium carbonate–bicarbonate mixture (1:10?w/w) yielded the RE oxide cake which constitute about 1.2% (w/w). The produced rare earth oxide cake contains traces of various metal oxides, e.g., SrO, Na₂O, etc. in addition to rare earth oxides. Attempts to determine quantitatively the constituents of the cake will be considered in future work.     

Proposed upgrading techniques for East Sibaiya phosphorites,Aswan area,southern Egypt

The Late Cretaceous (Campanian-Maastrichian), low-grade phosphorite sequence of East Sibaiya, Aswan area, which is known as the Duwi (phosphate) or Sibaiya Formation, is usually intercalated with marl, oyster limestone, and chert beds. These strata, which crop out in a generally east–west trending belt spanning the middle latitudes of Egypt, are phosphorite-rich sediments and of great economic importance. Representative samples were collected from the investigated area at several localities (e.g., Umm Tundubah-2 and Wadi El-Batur) and 200 kg of phosphate rocks was used for upgrading processes of low-grade phosphorites of the East Sibaiya area. The upgrading processes included two techniques: the first technique (gravity separation) comprises crushing, sieving, gravity separation by a shaking table, and magnetic separation. This technique raised the P₂O₅ in the head sample from 24.73 to 31.91% as a phosphate concentrate. The second technique (flotation technique) depends on certain flotation parameters such as pH, grain size, and phosphate collector dose (i.e., oleic acid). The flotation technique increased the P₂O₅% from 24.73 to 31.16% as a final product of the phosphate concentrate. These data were confirmed by X-ray fluorescence analyses of major elements.     

Two-stage reverse flotation process for removal of carbonates and silicates from phosphate ore using anionic and cationic collectors

Phosphate rock contains various gangue minerals including silicates and carbonates which need to be reduced in content in order to meet the requirements of the phosphate industry. Froth flotation has become an integral part of phosphate concentration process. In this study, double reverse flotation was applied to recover apatite from phosphate ore. H₃PO₄ and CaO were used as phosphate depressants, in acidic and alkaline conditions. Fatty acids and amines were added as carbonate and silicate collectors respectively. An experimental protocol devised to optimize the grade and recovery of phosphate using anionic–cationic method was found effective. Consequently, a required high quality of phosphate concentrate containing 30.1% P₂O₅ was obtained, with a recovery of 94%. X-ray diffraction and optical microscopy studies were performed to define the main minerals.     

Diagrams of electrochemical equilibria of the system copper-potassium ethylxanthate-water at 25°C

On the basis of thermodynamic calculations, diagrams of metastable electrochemical equilibria of the system copper-potassium ethylxanthate-water, at 25°C were constructed. The equilibria equations and diagrams for the total activity of [EtX^?] + [HEtX] + 2 [(EtX)₂] (equalling 10^?1, 10^?4 and 10^?7 mol/dm³), are presented. The practical conclusions resulting from these diagrams for the flotation of native copper are discussed.     

Microlaboratory study on magnetic,gravity and high-tension separation of hercynite and pleonaste from low-grade ilmenite concentrates

Some ilmenite concentrates obtained from oleate flotation of ilmenite ores from NE Poland contain less than the required 45% TiO₂ due to the presence of green spinels (hercynite and pleonaste). Such concentrates were further upgraded by different separation techniques. It was established at microlaboratory scale that magnetic, gravity and high-tension separations can provide qualified ilmenite concentrates with TiO₂ recovery in the order of 50 to 80%.     

Flotation behaviour of huntite (Mg3Ca(CO3)4) with anionic collectors

In this study, flotation behaviour and electrokinetic properties of huntite were determined in the presence of anionic collectors such as potassium oleate, and Flotinor FS-2. The effects of common ions that are expected in flotation pulps such as Mg²⁺ and Ca²⁺ were studied. The influences of pH and modifying reagents like sodium silicate (Na₂SiO₃·5H₂O) and carboxymethyl cellulose (CMC) on surface properties of huntite were also examined. Electrokinetic measurements showed that isoelectric point (i.e.p.) of huntite was around 8.0. Surface charge of huntite becomes positive in the presence of 10⁻² M of MgCl₂ and CaCl₂ at all pH values. According to micro-flotation test results, potassium oleate is more efficient compared with Flotinor FS-2. Mg²⁺ and Ca²⁺ ions have detrimental effect on the floatability of huntite. Na₂SiO₃ was seen to have no significant effect on the floatability of huntite. Flotation experiments with CMC indicated that it acts as a depressant on huntite.