The development of a sulfidisation technique to restore the flotation of oxidised pentlandite

The sulfidisation of a thermally oxidised Nkomati massive sulfide ore was successful in restoring the flotation recoveries of the oxidised sulfide minerals including pentlandite. As a single mineral, however, oxidised pentlandite was not readily sulfidised and XPS analyses failed to identify any sulfidised surface species. A sulfidisation mechanism for oxidised pentlandite is proposed whereby iron and copper species from the oxidised Nkomati mineral surfaces form sulfides on the oxidised pentlandite surfaces, either through a precipitation process or the conversion of adsorbed iron and copper hydroxides, and subsequently confer floatability. To evaluate this proposal, the role of base-metal ions during the sulfidisation process was studied for oxidised pentlandite as a single mineral. The addition of iron during sulfidisation was found to be effective in improving the subsequent flotation recovery and was a maximum after sulfidisation at a sulfide ion electrode potential of −650 mV (−315 SHE). While this corresponded to that found for pentlandite during the sulfidisation of oxidised Nkomati ores, the sulfidised pentlandite floated more slowly as a single mineral. Iron was more effective than copper, most likely due to a significantly higher concentration. The significance of the sulfidisation potential at lower hydrosulfide concentrations is related to hydrosulfide ion availability while at higher sulfidisation potentials, it is associated with the depressive effect of adsorbed base-metal hydrosulfide ion complexes. A technique is presented whereby base-metal ions additions during sulfidisation improved the flotation recovery of a difficult to sulfidise mineral species, such as oxidised pentlandite.     

Effect of Leptospirillum ferrooxidans on the flotation kinetics of sulphide ores

The objective of this work was to determine the effect of Leptospirillum ferrooxidans on the floatability of chalcopyrite, sphalerite, and pyrrhotite by using xanthate as a collector. The tests were carried out in the absence and presence of bacteria in relation to the type of ore and contact time with bacteria. The results indicate that the chalcopyrite flotation rate significantly increased in the presence of L. ferrooxidans due to the formation of hydrophobic species. The bacteria function as a weak depressant for pyrrhotite after a conditioning time ?60 min. The behaviour of sphalerite remains without changes due to its low susceptibility to oxidation. It was concluded that L. ferrooxidans brings about superficial changes mainly due to the oxidation of minerals.     

Pyrrhotite oxidation and its influence on alkaline amine flotation

In the production of ground calcium carbonate (GCC) for the paper industry, pyrrhotite dramatically reduces the GCC brightness and the removal of pyrrhotite through froth flotation is essential. The present study aims to study the effect of pyrrhotite oxidation on flotation recoveries during typical GCC flotation (i.e. alkaline pH, CaCO₃ saturation, amine collector). EDTA extraction and measurements of Eh (redox potential) showed a significant difference in pyrrhotite surface oxidation state when comparing exposure times of 40 and 60 min, the latter being significantly more oxidised. Microflotation results show that when pyrrhotite is exposed to air extending for more than 5 min at pH 8, the recoveries drop significantly. At higher pH, recoveries were generally low at all exposure times tested. Flotation recoveries showed strong correlation with zeta potential measurements. Bench scale flotation experiments on a sulphide bearing marble, confirmed that pyrrhotite oxidation significantly lower the GCC quality at low collector concentrations. By increasing the amine concentration, the flotation performance became independent of pH and exposure time.     

Geometallurgical characterisation of Merensky Reef and UG2 at the Marikana Mine,Bushveld Complex,South Africa

The influence of mineralogy on the milling performance and the flotation-based recovery of Au, the platinum group elements (PGE), Co, Cr, Cu, Ni and S was investigated for three samples of the Merensky Reef (BK, RPM and WP facies types) and one sample of the UG2 at the Marikana mine, using a mineral liberation analyser (MLA). The samples differ in their milling behaviour in that a range of milling times are required in order to produce a grind of 60% passing 75 μm. This is primarily controlled by the abundance of plagioclase, orthopyroxene and chromite.The size distribution of the base metal sulfides (BMS; pyrrhotite, pentlandite and chalcopyrite) is similar for the three samples of Merensky Reef, and is significantly coarser than for BMS in the sample of the UG2. Upon milling to 60% passing 75 μm, the best BMS liberation is achieved for the BK facies type of Merensky Reef, relative to RPM and WP. The degree of BMS liberation in the sample of UG2 is lower than that for samples of the Merensky Reef.Cumulative mass pull during flotation is higher for the sample of the WP facies of Merensky Reef than for the rest of the samples examined. This is due to the higher abundance of orthopyroxene in this sample, which is known to be naturally floating, and which reports to concentrate.A high flotation-based recovery of PGE, Cu and S is observed for all four samples, with the highest recovery associated with the sample of the BK facies type of Merensky Reef. Ni recoveries are generally poor, suggesting that Ni is associated with gangue minerals, in addition to that in pentlandite.Of the three facies types of Merensky Reef examined, the overall characteristics of the BK facies type i.e. a high PGE grade, low abundance of orthopyroxene, a high abundance of BMS, and a higher degree of liberation of the BMS on milling of the ore, represent the most favourable set of characteristics for the efficient recovery of PGE. It is therefore the best quality ore of the three samples of Merensky Reef examined.     

Adsorption of copper sulphate on PGM-bearing ores and its influence on froth stability and flotation kinetics

Copper sulphate is used as an activator in the flotation of base metal sulphides as it promotes the interaction of collector molecules with mineral surfaces. It has been used as an activator in certain platinum group mineral (PGM) flotation operations in South Africa although the mechanisms by which improvements in flotation performance are achieved are not well understood. Some investigations have suggested these changes in flotation performance are due to changes in the froth phase rather than activation of minerals by true flotation in the pulp zone. In the present study, the effect of copper sulphate on froth stability was investigated on two PGM containing ores, namely Merensky and UG2 (Upper Group 2) ores from the Bushveld Complex of South Africa. Froth stability tests were conducted using a non-overflowing froth stability column. Zeta potential tests and ethylenediaminetetraacetic acid (EDTA) tests were used to confirm the adsorption of reagents onto pure minerals commonly found in the two ores. The results of full-scale UG2 concentrator on/off copper sulphate tests are also presented. The UG2 ore showed a substantial decrease in froth stability in the order of reagent addition: no reagents > copper > xanthate > copper + xanthate, while Merensky ore showed a slight decrease. It was shown through zeta potential measurements that copper species were to be found on plagioclase, chromite, talc and pyrrhotite surfaces and through EDTA extraction that this copper was in the form of almost equal amounts of Cu(OH)₂ and chemically reacted copper ions on the Merensky and UG2 ore surfaces. In certain cases, the presence of copper sulphate and xanthate substantially increased the recovery, and therefore the implied hydrophobicity, of pure minerals in a frothless microflotation device. It was, therefore, proposed that increases in hydrophobicity beyond an optimum contact angle for froth stability, were the cause of instabilities in the froth phase and these were found to impact grade and recovery in a full-scale concentrator. Differences in the extent of froth phase effects between the different ores can be attributed to differences in mineralogy.     

Influence of mineralogy and ore texture on pentlandite flotation at the Nkomati nickel mine,South Africa

The influence of ore mineralogy and ore texture on flotation response was studied for 29 samples from the main mineralised zone at Pit 3 of the Nkomati Ni mine, through laboratory scale flotation testing, laboratory assay, and mineral liberation analyser examination of the ore and the concentrates. The individual sample flotation responses vary widely in terms of Ni grade, and cumulative Ni recovery. It is demonstrated that this is a complex function of ore mineralogy and ore texture. Chalcopyrite is the first sulphide to float, followed by pentlandite and finally pyrrhotite, in ore samples with dominant chalcopyrite, or where pentlandite, pyrrhotite and chalcopyrite occur in equal abundance. However in samples with a high ratio of pyrrhotite to pentlandite and chalcopyrite, pyrrhotite floats earlier than expected, reports to concentrate over the entire flotation period, and depress and extend the flotation of pentlandite over the flotation interval with no clear peak of Ni recovery during flotation. Primary silicates (e.g. olivine and pyroxene) and alteration-related minerals (talc, tremolite and chlorite) are naturally floating, and hence affect the flotation of pentlandite in a similar manner to that of pyrrhotite. The most problematic ore at Nkomati in terms of Ni recovery is characterised by fine disseminated and fine bleb- or net-texture sulphides, contain abundant olivine, pyroxene, amphibole, talc and tremolite, and include abundant metamorphism-related country rock xenoliths (with calc-silicate minerals such as diopside and tremolite).     

The effect of polysaccharides and polyacrylamides on the depression of talc and the flotation of sulphide minerals

Three polymers of varying chemistries and molecular weights have been used to depress talc in a model sulphide ore consisting of talc, pentlandite, and chalcopyrite. Flotation testing revealed a distinct hierarchy of effectiveness of the polymers in the depression of talc (HP Starch > Polymer-H > Dextrin-WY) but a different order of effectiveness for depression selectivity (Dextrin-WY > Polymer-H > HP Starch). Adsorption isotherm determinations for the gangue and valuable mineral phases indicate that the differences in selectivity are partially due to adsorbed amount variations, with Dextrin-WY adsorbing marginally less on the sulphides than Polymer-H. Contact angle measurements on packed particle beds have revealed that the polymers have differing abilities to reduce the contact angle of talc and the sulphide minerals. An explanation of the observed trends is given in terms of the binding mechanisms of the polymers on the talc and the sulphide mineral phases. Order of addition effects between collector and depressant were also investigated using adsorption isotherm determinations for the polymers (collector) in the presence of a pre-adsorbed layer of collector (polymer). Adsorption of the polymers was not affected by the presence of the collector and vice versa.     

Bioleaching of six nickel sulphide ores with differing mineralogies in stirred-tank reactors at 30 °C

A bioleaching study was conducted with six nickel sulphide ores from different geographical locations across Canada. Mineralogical and chemical examination revealed considerable variability between the samples, particularly in the silicate phases. The ores contain 0.3–1% nickel, primarily in pentlandite and secondarily in pyrrhotite. Copper is present primarily in chalcopyrite, and cobalt in pentlandite. The ores were subjected to the same crushing and grinding procedure, and bioleached under the same conditions for 3 weeks with a mixed culture of iron- and sulphur-oxidizing bacteria. Stirred-tank experiments with finely ground ore (−147 μm) at 30 °C were conducted to assess the effect of pH (2–5) and the impact of the bacteria. Nickel extraction from pentlandite and pyrrhotite during bioleaching at pH 2 and 3 was generally good (49–86% after 3 weeks), and cobalt extraction tracked nickel extraction over most conditions. All six ores showed a similar response to a change in pH; an increase in pH from 2 to 3 resulted in approximately the same nickel and cobalt extraction (within statistical error), and a statistically significant reduction in sulphuric acid consumption, dissolved iron, and magnesium extraction.     

Change in technological properties of technogenic sulfide-contained raw material in the storage process

Results of modeling of changes in technological properties of sulfide minerals (pyrrhotite, pentlandite, and chalcopyrite) present in mining-industrial wastes stored in a climate in which evaporation prevails over precipitation are reported. It is proved that flotation activity of chalcopyrite and, in particular, of pentlandite diminishes, pyrrhotite increases, and selectivity and flotation of combined sulfides with respect to other minerals increases insignificantly. The content of water-soluble salts, dominated by nickel sulfates, increases with storage time. Institute of Problems of Complex Development of Bowels, Russian Academy of Sciences, Moscow, Russia Translated from Fiziko-Tekhnicheskie Problemy Razrabotki Poleznykh Iskopaemykh, No. 3, pp. 108–114, May–June, 2000.     

Flotation of pyrrhotite and pyrite in saturated CaCO3 solution using a quaternary amine collector

In the production of ground calcium carbonate (GCC) for the paper industry, any colouring contaminants must be removed during processing (usually flotation) to achieve sufficient GCC brightness. Flotation of GCC feedstock is characterised by the presence of both particulate and dissolved calcite (CaCO₃), an alkaline pH and the use of amine collectors. This paper investigates the possibility of removing pyrrhotite and pyrite, under these conditions. Microflotation results show that the recoveries are highly dependent on pH. Pyrrhotite recovery decreases dramatically when going from pH 8 to pH 10. Pyrite display high recoveries at pH 8 and 9, and lower recoveries at higher pH. Recoveries were dependent on conditioning time. Pyrite floats considerably faster than pyrrhotite at all pH levels investigated, whereas pyrrhotite seems to be slow floating and could benefit from prolonged flotation times. Compared to the saturated CaCO₃ system, flotation recoveries decreased when using deionised water or CaCl₂ solution. The flotation results could not be fully explained by zeta potential. Bench scale flotation experiments on a sulphide bearing marble, confirmed the microflotation results at low collector concentrations (i.e. 200 g/t). By increasing the amine concentration, the flotation performance became independent of pH.     

某铜镍多金属矿石工艺矿物学研究

某铜镍多金属矿床赋存于蚀变超基性岩镍矿床中,伴生铜、钴及贵金属铂、钯、金、银.通过显微镜观察,采用X射线衍射仪、扫描电镜能谱仪及矿物自动检测仪等分析技术,对该矿床的物质组成、目的矿物嵌布特征、有价元素平衡分配等工艺矿物学参数进行了系统的研究.结果表明,矿石中目的矿物主要为镍黄铁矿和黄铜矿,并含砷铂矿、锑铂矿、自然金、银...     

Interfacial properties of liquid sulfur in the pressure leaching of nickel concentrate

The effects of sulfur dispersing agents in the oxygen pressure leaching of nickel concentrate at medium temperature were investigated by interfacial studies. Liquid sulfur–aqueous solution interfacial tensions and liquid sulfur–sulfide mineral contact angles were measured at 140 °C, 690 kPa overpressure by nitrogen. The effects of sulfur dispersing agents including lignosulfonate, Quebracho, o-phenylenediamine (OPD), and humic acid were evaluated by the calculation of the work of adhesion in the liquid sulfur–sulfide mineral–aqueous solution systems. It was found that the sulfide mineral surface is sulfophobic at pH from 4.1 to 4.5 due to the hydrolysis of nickel (II) ions to nickel hydroxide and the deposition of nickel hydroxide on the mineral surface. These findings apply to four different sulfide mineral systems, including pentlandite, nickeliferous pyrrhotite, pyrrhotite, and chalcopyrite. Lignosulfonate, Quebracho, and humic acid were found to significantly reduce the work of adhesion indicating they should be effective sulfur dispersing agents. OPD is ineffective in changing the work of adhesion of sulfur on the mineral sulfides indicating that it is not a good candidate for sulfur dispersion.     

Sorption of mercaptanbenzothiazol and dithiophosphate on Pt-Cu-Ni minerals at flotation process

The test data on sorption and flotation properties of reagent-collectors: dithiophosphate (DTP) and mercaptanbenzothiazol (MBT) relative to Pt-Cu-Ni minerals (chalcopyrite, pentlandite, and pyrrhotite) are presented. It was established by UV-spectroscopy analysis that MBT sorption on the study mineral specimens amounted to 62 % for chalcopyrite, and 15–30 % for pyrrhotites, 88 % for pentlanditepyrrhotite specimen to the initial reagent quantity. Adsorption of isopropyl and isobutyldithiophosphate on pentlandite-pyrrhotite was 59% and 82 %, respectively, along with none actual adsorption of these reagents on pyrrhotite.. It is shown that the selective sorption of MBT and DTP proves the expediency of applying the above complex-forming reagent mixture as selective collectors for platinum-bearing sulfide minerals at Pt-Cu-Ni ore flotation. __________ Translated from Fiziko-Tekhnicheskie Problemy Razrabotki Poleznykh Iskopaemykh, No. 6, pp. 129–134, November–December, 2007.     

Selective separation of fine albite from feldspathic slime containing colored minerals (Fe-Min) by batch scale dissolved air flotation (DAF)

In this study, the separation of feldspar minerals (albite) from slimes containing feldspar and iron containing minerals (Fe-Min) was studied using dissolved air flotation (DAF) technique whereby bubbles less than 100 μm in size are produced. Before the flotation experiments with slimes, single flotation experiments with albite and Fe-Min were carried out using DAF in order to obtain optimum flotation conditions for the selective separation of feldspar from the slimes. Flotation experiments were performed with anionic collectors; BD-15 (commercial collector) and Na-oleat. The two methods of reagent conditioning were tested on the flotation performance; traditional conditioning and charged bubble technique. In addition, the effect of pH, flotation time, rising time, and drainage time which influence the selective separation in the DAF system were studied in detail. Overall, the flotation results indicated that the separation of albite from Fe-Min can be achieved with DAF at 5 min of rising time and 5 min of drainage time. Interestingly, these results also showed that the conditioning of the particles with the charged bubbles increased the flotation recovery of Fe-Min compared to the traditional conditioning. Furthermore, the flotation tests with the feldspathic slime sample were carried out under the optimum conditions obtained from the systematic studies using the single minerals. The charged bubble technique produced an albite concentrate assaying 0.33% Fe₂O₃ + TiO₂ and 11.07% Na₂O + K₂O from a slime feed consisting of 1.06% Fe₂O₃ + TiO₂ and 10.36% Na₂O + K₂O.     

A comparison of pyrrhotite rejection and passivation in two nickel ores

The non-stoichiometric sulfide mineral pyrrhotite (Fe1-xS) occurs almost ubiquitously inter-grown with the principal nickel mineral, pentlandite ((Fe,Ni)9S8). During Ni processing, pyrrhotite is generally rejected to the tailings stream by flotation to produce a low tonnage, high grade (Ni) smelter feed and reduce SO₂ emissions. In this study, the effect of different pyrrhotite flotation rejection strategies (artificial oxidation and TETA: SMBS addition) are evaluated on a magnetic (Ore A) and non-magnetic (Ore B) pyrrhotite ore to determine if either may effectively depress and potentially passivate the pyrrhotite surface during flotation to produce benign tailings without compromising pentlandite recovery. For both ores, the best pyrrhotite rejection (pentlandite/pyrrhotite recovery) was obtained using TETA: SMBS. Differences in the flotation performance of the two ores are considered more a function of BMS content, liberation and ore handling rather than a difference in sulfide passivation from the inherent pyrrhotite mineralogy (magnetic vs non-magnetic pyrrhotite). Pyrrhotite passivation could possibly provide a means of rendering the tailings non-reactive and thus mitigate acid rock drainage (ARD) formation.     

某复合型铜钴矿合理选矿工艺及硫化作用机理研究

以NaHS作硫化剂、水玻璃和CMC作组合抑制剂、Y-89作捕收剂、MIBC作起泡剂, 采用预先浮选脱泥、NaHS诱导同步浮选的选别工艺流程对某复合型铜钴矿进行了选矿工艺研究。原矿中铜、钴品位分别为1.14%、0.18%时, 闭路精矿产品中, 铜的品位达到24.78%, 回收率为72.66%; 钴的品位为2.75%, 回收率达到51.10%。硫化剂诱导浮选氧化矿的作用机理分析表明, 硫化剂起作用的主要组分是HS^-, NaHS与Na₂S对浮选回收铜都表现出了很好的适应性, 但在相同用量条件下, NaHS由于能够在钴矿物表面形成更加稳定的金属硫化物膜, 因而对钴的浮选回收效果更好。     

The role of the reagent suite in optimising pentlandite recoveries from the Merensky reef

The Platinum group elements (PGEs) in the Merensky Reef in the Bushveld Complex are strongly associated with the sulphide minerals, particularly pentlandite which contains significant amounts of PGEs such as palladium and rhodium in solid solution. The various operations use different reagent combinations in the flotation process to recover PGEs from the Merensky Reef which include primary and secondary collectors, copper sulphate (CuSO₄) as an activator and depressants, either carboxymethylcellulose (CMC) or modified guar gum (guar). This study examines the effect on two different ores, A and B, using dibutyl dithiophosphate (DTP) as a secondary collector in combination with sodium isobutyl xanthate (SIBX) as primary collector, the point of collector addition (mill or cell), the use of CuSO₄ addition, sequence of CuSO₄ addition and the effect of depressant type on the floatability of pentlandite, and shows that pentlandite floatability in contrast to that of chalcopyrite is reduced when reagents were added to the cell rather than the mill together with CMC. This effect was not observed with the use of guar gum as the depressant.     

Improved statistical methods applied to surface chemistry in minerals flotation

Diagnosis of the surface chemical factors playing a part in flotation separation of a value sulfide phase requires measurement of the hydrophobic and hydrophilic species that are statistically different between the concentrate and tail streams. Statistical methods, based on the monolayer-sensitive time of flight secondary ion mass spectrometry (ToF-SIMS) technique, have been developed towards this ultimate aim by measuring hydrophobic species (collector ions, dimers and metal complexes, polysulfides) as well as hydrophobic metal ions, precipitates and added depressant species. Reliable identification of specific mineral particles is central to this statistical analysis. A chalcopyrite/pyrite/sphalerite mineral mixture conditioned at pH9 for 20 min to study transfer of Cu from chalcopyrite via solution to the other two mineral surfaces, since this mechanism can be responsible for their inadvertent flotation in copper recovery, showed no statistical difference in the copper intensities on pyrite and sphalerite (selected from Fe and Zn images) after this conditioning. Principal component analysis (PCA) identifies combinations of factors strongly correlated (positively or negatively) in images or spectra from sets of data. In images, PCA selects these correlations from the mass spectra recorded at each of 256 × 256 pixels in a selected area of particles. In the image mode, PCA has proved to be a much better method of selecting particles by mineral phase with clearer definition of particle boundaries due to multi-variable recognition. It has clearly separated a statistical difference in copper intensities between the sphalerite and pyrite phases.The PCA method has been applied to concentrate and tails samples collected from the Inco Matte Concentrator demonstrating extensive CuOH and NiOH transfer between the chalcocite (Cc) and heazlewoodite (Hz) minerals. Statistical differences illustrate the important discriminating depressant action of NiOH in flotation despite the activation of Hz by Cu transfer. The adsorption of the collector at specifically-identified Cu sites has been elucidated by the study. Importantly, the statistical analysis has been able to confirm some mechanisms and deny others proposed to control recovery and selectivity giving more focus on the control factors.     

Flotation of rare earth minerals from silicate–hematite ore using tall oil fatty acid collector

The flotation of rare earth (RE) minerals (i.e. xenotime, monazite-(Nd), RE carbonate mineral) from an ore consisting mainly of silicate minerals (i.e. primary silicate minerals and nontronite clay) and hematite was investigated using tall oil fatty acids (Aero 704, Sylfat FA2) as collector. The RE minerals are enriched with Fe. The effects of tall oil fatty acid dosage, pH, temperature, and conventional depressants (sodium lignin sulfonate, sodium metasilicate, sodium fluoride, sodium metasilicate and sodium fluoride, and soluble starch) were determined at grinding size of P80 = 63 μm. At this grinding size, the grain size of the RE minerals ranges from 2 to 40 μm, percentage liberation is 9–22%, and percentage association with nontronite and quartz is 30–35%. Results indicated that Sylfat FA2 at 22450 g/t concentration was the more efficient tall oil fatty acid collector at natural pH (pH 7) to basic pH (pH 10.0–11.5). Flotation at the room temperature (25 °C) gave higher selectivity than 40 °C temperature flotation. The results on the effect of depressants showed similar selectivity curves against the gangues SiO₂, Al₂O₃, and Fe₂O₃ suggesting that the chemical selectivity of the depressants has been limited by the incomplete liberation of the RE minerals in the feed sample. High recoveries at 76–84% (Y + Nd + Ce)₂O₃ but still low (Y + Nd + Ce)₂O₃ grade at 2.1% in the froth were obtained at flotation conditions of 63 μm, 25 °C, pH 10.5, 1,875 g/ton sodium metasilicate and 525 g/ton sodium fluoride or 250 g/ton soluble starch as depressant for the silicates and hematite, and 22,450 g/t Sylfat FA2 as collector for the RE minerals (initial (Y + Nd + Ce)₂O₃ feed grade = 0.77%). The recoveries of gangue SiO₂, Al₂O₃, and Fe₂O₃ in the froth were low at 25–30%, 30–37%, and 30–36%, respectively. The mineralogical analysis of a high grade froth and its corresponding tailing product showed that the RE minerals have been concentrated in the froth while the primary silicate minerals and hematite have been relatively concentrated in the tailing. However, the clay minerals, primary silicate minerals, and hematite still occupy the bulk content of the froth. This suggests that incomplete liberation of the RE minerals led to the poor grade result, supporting likewise the selectivity curve results by the different depressants. This study showed that liberation is important in achieving selective separation.     

A modification in the flotation process of a calcareous–siliceous phosphorite that might improve the process economics

The amenability of a low-grade Egyptian phosphorite to flotation for separation of both calcareous and siliceous gangue minerals by just pH control was investigated. The ore, assaying 19.39% P₂O₅, 16.1% L.O.I. and 12.41% A.I. is mainly composed of francolite and hydroxy apatite minerals consolidated into three different phosphatic varieties according to texture and origin, i.e. coarse phospho-chem, sharp-edged phospho-clast and fine cementing phospho-mud. This was endorsed by microscopic investigation of thin sections. X-ray diffraction analysis of the ore sample showed that the main gangue minerals are calcite and quartz with minor dolomite and some gypsum.Anionic flotation of calcite, under pH4.5, was successfully conducted on the −0.25 + 0.074 mm phospho-chem fraction without any use of phosphate depressants. This was followed by direct flotation of phosphate after raising the pH to 9. Mechanical cleaning of the phospho-concentrate was carried out, without any addition of the collector to get rid of the entrained silica. About 3 kg/t of oleic acid was required for the whole process which was added step-wise 0.5 kg/t each except for the first step which was 1.0 kg/t to activate the flotation pulp. Phospho-concentrate assaying 30.54% P₂O₅, 8.7% L.O.I. and 5.76% A.I. with a P₂O₅ recovery of 64.34% was finally obtained without the use of expensive depressants, e.g. phosphoric acid or sodium silicate.A trial to explain the results in view of others’ findings and in terms of the ore mineralogical characteristics was shown.