Flotation of Low-grade Phosphate Ore

Advanced Powder Technol., Vol. 15, No. 4, pp. 421–433 (2004) © VSP and Society of Powder Technology, Japan 2004. Also available online - www.vsppub.com www.vsppub.com

Original paper

Flotation of low-grade phosphate ore F. HERNÁINZ ∗ , M. CALERO and G. BLÁZQUEZ Departamento de Ingeniería Química, Facultad de Ciencias, Universidad de Granada, 18071 Granada, Spain

Received 22 July 2003; accepted 26 September 2003 Abstract —A

study has been conducted on the flotation concentration of sedimentary phosphorite from Sierra Espuña, in the province of Murcia (Spain), using a Denver D-12 mechanical subaeration cell. First, the mineral was pretreated through grinding, sorting and discarding slimes. The effect on flotation of pH, collector concentration, particle size and flotation time was then analyzed. The results achieved achie ved show show that that a neutra neutrall or slightly slightly alkal alkaline ine pH and short short flotation flotation times promot promotee the conce concentrat ntration ion of sedimentary phosphorite. Keywords: Flotation; phosphate ore; oleic acid; particle size; mechanical cell.

1. INTRODUCTION

Phosphate materials (phosphates) occur naturally both in water (rivers, lakes, seas) and,, in higher and higher concen concentra tratio tions ns,, in soil. soil. It Itss us usee as a ra raw w mat materi erial al in the fertili fertilize zerr industry indu stry (sup (superph erphosph osphate ates) s) is the app applica lication tion of grea greatest test inte interest rest for soil soil-bo -borne rne phosphatee material phosphat materials. s. Phosphates can be grouped, according to types, in three major groups [1]: • Marine sedimentary phosphorites. These are the most important in both amount and volume. Their P 2O5 content may be over 30%. phosphorites. The apatite in these minerals comes from igneous–alkaline igneous–alkaline • Pluton phosphorites. complex comp lexes. es. Thes Thesee are high higher er in numb number er,, bu butt poorer and smaller smaller than mari marine ne sedimentary phosphorites. • Guano phosphorites. These are a result of reactions between marine bird and bat excrements excreme nts with w ith limestone rocks, giving rise to calcium phosphate. Phosphorus concentrations in phosphates can be expressed as their P 2O5 content, although Ca3 (PO4 )2 content (BPL) can also be used. The equivalence between the ∗

To whom correspondence should be addressed. E-mail: [email protected]

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two is: 1% P 2O5 = 2.185% Ca3 (PO4)2 . Currently, around 80% extraction rates are being achieved for sedimentary phosphates, 18% for igneous types and 2% for guano phosphates. Studying the mineralogy of sedminary phosphates poses considerable difficulties as these substances are rarely crystallized and also contain amorphous phosphates mixed with calcareous impurities. Its structural formula is generally accepted as being [2]: Ca 10 [(PO4 )6−x (CO3F)x ]F−2 , although part of the calcium (Ca 2+) may be replaced by sodium (Na + ) and magnesium (Mg 2+ ), while part of the fluor (F −) may be replaced by OH − . The effects of weathering, metamorphism and geological time may also give rise to systematic changes in the composition of the phosphorites. The industrial requirements for superphosphates demands a P 2 O5 content in the phosphoric ore close to or above 30% (approximately 60% in BPL) so as to lower sulfuric acid consumption [2]. The presence of free carbonates in phosphoric rock leads to considerably higher sulfuric acid requirements in wet treatment of this raw material, for obtaining both superphosphates and phosphoric acid. Apart from its role as an air pollutant, the CO 2 produced by the acid attack causes foaming that gives rise to the formation of small gypsum crystals that reduce the quality of the final phosphoric acid [3]. In the light of the above, low-grade phosphorites (P 2 O5 below 20–25%) must be concentrated prior to their industrial enrichment. The techniques generally applied for this purpose are: calcination , when the ore has low or average carbonate content, and flotation, when the ore has a high carbonate content and siliceous gangue. Both techniques have been studied in depth over the last few decades [4–10]. Calcination methods, however, do not generally give optimum results when the concentrated mineral is intended as the basis for superphosphate production, as the nodules formed during calcination may hinder acid attack (Avila Roson, pers. commun.). Recently, Sadeddin and Abu-Eishad [11], and Abu-Eishad et al. [12] proposed diluted solutions of acetic acid as a solubilizing agent for the carbonates in the acid leaching of the ore. For the time being, the procedure is restricted to laboratory experientation with poor prospects for industrial-scale implementation due to the costs of the reagents involved. From the above, it is clear that flotation is the most widely used treatment for the concentration of low-grade phosphorites, although there are numerous variations in techniques given the enormous variations in mineral composition from different deposits, especially with regard to carbonated and siliceous ores. It is this kind of ore that usually guides the use of flotation reagents and of one or more flotation stages, as well as setting operational parameters [2]. The P2 O5 content of the sedimentary phosphorite of Sierra Espuña is about 5% and is considered a low-grade ore. The concentration the low-grade phosphate by flotation is complicated because the fattyacid collectors used to float phosphate minerals also float carbonate minerals [13]. In recent years, several useful processes

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have been proposed for the concentration of low-grade phosphate ores by froth flotation using both anionic- and cationic-type collectors [14]. In this paper, the flotation of sedimentary phosphorite from Sierra Espuña in the province of Murcia (Spain) in a semi-industrial subaeration cell (Denver D-12) was studied to analyze the effect of the operational parameters which have the greatest influence on the concentration of the phosphate fraction in the mineral. The collector used in this study is oleic acid (anionic type).

2. MATERIALS AND METHOD

2.1. Mineral

Sedimentary phosphorites from a deposit located in the north of Sierra Espuña (province of Murcia), in the area known as Prado Mayor on the left slope of the Malvariche ravine [15], were used for this study. Following extraction of the mineral from the deposit in 10 × 10 cm cores taken along a stretch of 10 m, the mineral was mixed in the laboratory to achieve homogenization so as to obtain a reliable, representative sample of the deposit. The mineral was then ground in a three-floating hammer Culatti micromill and sorted according to grain size in a high-vibration machine equipped with a set of fine ASTM sieves, with the selection of appropriate sizes. Table 1 provides the equivalents for the sieves used during the study together with mineral size. Table 1.

Sieve ASTM (mesh)

Mineral size (μm)

20 30 40 50 60 70 80 100 120 140 160 200 270 325 400 450 500

840 590 420 297 250 210 177 149 125 105 88 74 53 44 37 30 25

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2.2. Flotation agents

The chemicals used for this study were as follows: • Collector . Chemically pure oleic acid supplied by Carlo Erba. This agent was diluted in distilled water just prior to the experiments to avoid any possible timeinduced alterations. • pH-modifying agents . Merck sodium hydroxide and Probus hydrochloric acid. Both were used at 4 N concentration to easily achieve the desired pH; pH was controlled by Crison 2001 pH meter. 2.3. Experimental method and conditions

A Denver-type free-flow mechanical subaeration cell, model D-12 was used as the flotation cell, as described in detail in [16]. The pre-prepared mineral was added to the 1.25-l flotation cell. The solution of surfactants was then added at the desired pH and the mixture was shaken for 5 min, i.e. sufficient conditioning time to ensure that the mineral was completely covered by the collector. Shaking was performed by the Denver cell itself at 1000 r.p.m. with the air valve closed. Flotation commenced upon completion of conditioning time. The foams collected during flotation time were vacuum-filtered, heat stove dried to a constant weight for later analysis. 2.4. Analytical method

The official method for analysis described in Annex II of the Order dated 30 November 1976 (published in the Spanish Official Gazette on 30 January 1977) was used to determine the floated mineral concentration, expressed in % P2O5total . The technique comprises a sulfonitric attack on the phosphoric acid to solublize the phosphoric acid and precipitate the molybdenic acid and quinolein, and collection and weighing of the yellow quinolein phosphomolybdate precipitate, from which P2 O5 content can be deduced. 2.4.1. P2 O5 analysis.

Calcium analysis was performed by atomic absorption spectrophotometry, using a Perkin-Elmer model 3100. The sample solution was prepared by alkaline fusion with lithium metaborate. A multi-element hollow cathode lamp supplied by Perkin-Elmer was used for measurement, together with an air-acetylene flame as required for these elements. 2.4.2. CaCO3 analysis.

The analysis of silicon (expressed as SiO 2 ) was performed by visible region absorption spectrophotometry using a Spectronic model 2000. 2.4.3. SiO2 analysis.


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3. RESULTS AND DISCUSSION

3.1. Mineral pre-treatment

Through previous tests, P 2O5 content in the mineral to be floated was proven to depend, to a considerable extent, on mineral size. This fact has already been addressed by Hegner and Pacl [17] when they observed that the P 2 O5 content in certain sedimentary phosphates ranged from 30 and 26% for coarse ( >0.16 mm) and fine ( <0.1 mm) fractions. In this paper, the treatment was as follows. The mineral, sampled by surface cores from the deposit, was quartered successively to obtain a weight of 2 kg; then sorted according to grain size in a digital electromagnetic sieve. Only grain sizes between 20 and 60 ASTM mesh were selected as these hold the highest P 2 O5 content (6–10%) and represent 40.39% of the original mineral. Afterwards, the mineral was ground and sorted, selecting only grain sizes between 20 and 50 ASTM mesh. These were then subjected to a second grinding and were sieved so that the original mineral’s P 2 O5 content amounted to approximately 10% in a wide range of sizes (from 60 to <200 ASTM mesh) which, in some cases, may even increase P 2 O5 content 4-fold. It is also well established [18–20] that the presence of ultrafine particles, known as slimes, has a very detrimental effect on the flotation of solids, as they contaminate foams, raise the consumption of flotation reagents, reduce the speed of the process and cause an anarchic response in mineral separation using this technique. For this reason, the mineral was treated to discard slimes by lavage, i.e. the ground mineral is washed with water over a 450 ASTM mesh sieve and then the rejects are sieved, selecting only sizes under 100 ASTM mesh. To verify the soundness of this procedure, laser grain size analysis was performed on the particles prior to and following discarding slimes; the results are shown in Fig. 1. As can be seen, the presence of extremely fine particles (slimes) is considerably higher in unwashed mineral. Furthermore, the particles below 12–14 μm are only found in unwashed mineral. 3.2. Mineral flotation

The effect of pH on the flotation of sedimentary phosphates was analyzed using oleic acid at a concentration of 1000 g/t as the collector, with washed mineral with a particle size below 100 ASTM mesh, 5 min conditioning time, modified pH between 7 and 11. The results obtained are shown in Fig. 2. The flotation recovery for the mineral is greater with a neutral or moderately alkaline pH; recovery is severely reduced in clearly alkaline flotation baths. Hence, at pH 8 over 50% of the mineral is recovered, while at pH 11 recovery drops to approximately 30%. Singh et al. [8], Hanumantha Rao et al. [21], Gong et al. [22] and Espinosa [23] found similar results when working with different phosphate minerals. 3.2.1. Influence of pH.

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Figure 1.

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Laser grain size analysis of the washed and unwashed particles.

As for P2 O5 , CaCO3 and SiO2 contents both in concentrate and tailing, results show that CaCO 3 generally appears in the concentrate, while P 2O5 is concentrated in the tailing. This gives rise to CaCO 3 contents between 62 and 78% in concentrate, while tailing content amounts to between 24 and 36%. P 2O5 content is about 2–3% in the concentrate and about 6.5–8% in the tailing. There is little change in silica content in concentrate and tailing, with content ranging from around 7% in concentrate to slightly above 8% in the tailing. Bearing in mind that the composition of the original mineral for this series of experiments was P 2 O5 5.56%, CaCO3 51.11% and SiO 2 7.42%, it can be deduced that flotation of the mineral leads to a significant concentration in calcite (CaCO 3) in the floated foams (in some cases, with an increase of over 25%), while the most valuable mineral is concentrated in the tailing, i.e. P 2 O5, with an 8% rise in proportion. As the aim is to raise the P 2O5 content, the operation performed is a reverse flotation given that the most valuable fraction is found in the tailing. This is a frequently used operation in the mining industry [18]. The distribution of P 2 O5 and CaCO3, in percentage terms, in both the concentrate and tailing has been calculated taking into account the recovery rates for total mineral, P2O5 and CaCO3 in the concentrate and tailing as well as content of both in the original mineral. The results are shown in Fig. 3. The distribution of SiO 2 in the ccentrate and tailing has not been provided given that SiO 2 content in both fractions is similar (7–8%), as mentioned above. The results shown in Fig. 3 also include the percentage of the content for each component. Bearing in mind both distribution and content, 75% of total P 2 O5 remains in the tailing at pH 7 with a concentration of 7.82% and 85% at pH 10 with a concentration of 7.28%. Concentration drops to 6.64% only at pH 11, although over 85% of the original mineral present is recovered. The P 2O5 content in concentrate is


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similar in all cases (ranging between 2 and 3%), and its distribution varies between 15 and 30% in the pH range studied. The best results for CaCO 3 are obtained in the concentrate; in the pH range between 7 and 10, around 60% is recovered with concentrations ranging between 62 and 78%. From these results, it is clear that separation of P 2O5 and CaCO3 through floation is best performed in flotation baths with a neutral or slightly alkaline pH. As the Sierra Espuña phosphates have the characteristic of yielding a natural pH of between 7 and 8 in the flotation bath, it was decided that the study would maintain this pH. This has enormous advantages for industrial applications as no reagents are required to modify pH and the possible surface attack of the mineral due to the presence of such reagents is also avoided. To study the effect on mineral recovery percentages of modifications in collector concentrations, a mineral size below 100 ASTM mesh and a flotation bath with the mineral’s natural pH were selected. Figure 4 shows the recovery percentage of the mineral according to oleic acid concentrations ranging between 250 and 2500 g/t. As the collector concentration rises, mineral recovery also increases, very rapidly initially (250– 1000 g/t) up to values of 60% at oleic acid concentrations above 1500 g/t. There is no further improvement in mineral flotation above this range. These results resemble those found by González Mantero [24] in a Hallimond tube using the same mineral, although with the Denver D-12 cell the recovery percentages are considerably higher. These results also agree with those reported by most researchers showing enhanced operations with an increase in collector concentration. Figure 5 provides the distribution of P 2O5 and CaCO3 between the concentrate and tailing, together with their content. As oleic acid is increased, there is a trend towards an even distribution between the concentrate and tailing for P 2 O5 , although concentration is much higher in the tailing. A similar effect occurs for CaCO 3, although concentration in the concentrate remains much higher than in the tailing. Given that the distribution of P 2 O5 and CaCO3 in concentrate and tailing is the reverse of their separation through flotation, the best results are achieved at oleic acid concentrations close to 1000 g/t as this yields CaCO 3 concentrations above 75% in the concentrate, with a distribution of over 80%, while P 2O5 is mostly found in the tailing (at around 70%) with concentrations above 7%. It can also be seen from Fig. 5 that very high collector concentrations, although they improve mineral flotation in all cases, are not appropriate for separating the different species present, as they all improve their flotation, which agrees with reports by other authors [25]. 3.2.2. Influence of oleic acid concentration.

Although in the preceding experiments the <100 ASTM mesh) is considered to be the most

3.2.3. Influence of particle size.

mineral size selected (i.e.

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Figure 2.

Flotation recovery of mineral versus pH with 1000 g/t oleic acid and particle size < 100 mesh.

Figure 3.

Flotation recovery of mineral versus oleic acid concentration at particle size < 100 mesh.

Figure 5.

Figure 4.

Distribution of P 2 O5 and CaCO3 , in percentage terms, in both the concentrate and tailing versus pH with 1000 g/t oleic acid and particle size < 100 mesh.

Distribution of P 2 O5 and CaCO3 , in percentage terms, in both the concentrate and tailing versus oleic acid concentration at particle size < 100 mesh.


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appropriate for flotation [18], a series of further experiments was also conducted with the <100 mesh size subdivided into four fractions, i.e. 100/160, 160/200, 200/270 and <270 ASTM mesh, so as to determine which fraction provides the best results. An oleic acid concentration of 1000 g/t was selected for these experiments. The results obtained are shown in Fig. 6, where mineral recovery is seen to improve at lower mineral sizes, down to a given size below which the recovery decreases. Accordingly, 200/270 mesh is the optimum size, giving recoveries approaching 65%, while 100/160 size only yields recoveries of nearly 50%. The results reported by González Mantero [24] in a Hallimond tube coincide entirely with regard to the most appropriate particle size and recovery percentages, which confirms that, in spite of its limitations, the Hallimond tube can estimate the influence of some of the parameters affecting mineral flotation quickly using only a small amount of mineral. The fact that this study also showed that, below a given solid particle size, there was no improvement in flotation basically agrees with Trahar’s findings [26] with sulfide minerals, with Lay and Bell [27] using saline minerals, with Bustamante and Warren [28] in carbon flotation, and with Hernáinz and Calero [29] in celestite ore. Figure 7 shows distribution rates between the concentrate and tailing for both P2 O5 and CaCO3 contents, respectively. The largest particle size of all used for P 2 O5 (100/160 mesh) provides the best distribution in the tailing, with recovery rates of almost 75% with a concentration of 8.67%. As size decreases, however, there is not only a decrease in mineral presence in the tailing, but content also decreases down to 6.41% for sizes below 270 mesh. These results are absolutely consistent, as mentioned earlier, with the fact that the grinding of the mineral concentrates the phosphate fraction in the larger size particles. Thus, together with the fact the P 2O5 concentrates predominantly in the tailing, the 100/160 mesh size yields the best results in the range of particle sizes used. For CaCO3 however, the best results are achieved at particle sizes of 160/200 and 200/270 mesh. The concentration of calcite is very similar, however, for most particle sizes and is only slightly lower for particle sizes below 270 ASTM mesh. From the above, it is clear that foams will concentrate CaCO 3 predominantly if the mineral is ground to particles sizes of 100–270 mesh, while P 2 O5 will mainly appear in the flotation cell residue. Two oleic acid concentrations, i.e. 1000 and 2000 g/t, together with four particle sizes, i.e. 160/200, 200/230, 230/270 and 270/325 ASTM mesh, were selected to analyze the influence of flotation time (t F ); flotation time was set at 1–10 min. Figure 8 shows the results obtained with 1000 g/t oleic acid, plotting the accumulated recovery percentage against flotation time for each particle size tested. As flotation time increases, the accumulated recovery percentage is also seen to rise, with the most marked effect occurring with 3.2.4. Influence of flotation time.

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Figure 6.

Flotation recovery of mineral versus particle size with 1000 g/t oleic acid.

Figure 7.

Flotation recovery of mineral versus flotation time at four particle sizes with 1000 g/t oleic acid.

Figure 9.

Figure 8.

Distribution of P 2 O5 and CaCO3 , in percentage terms, in both the concentrate and tailing versus particle size with 1000 g/t oleic acid.

Flotation recovery of mineral versus flotation time at four particle sizes with 2000 g/t oleic acid.

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Table 2.

[Oleic acid] = 1000 g/t Size (ASTM) (min) 160/200 200/230 [content (%)] 230/270 [content (%)] 270/325 [content (%)] P2 O5 CaCO3 SiO2 P2 O5 CaCO3 SiO2 P2 O5 CaCO3 SiO2 P2 O5 CaCO3 SiO2 t F

1.0 1.5 2.0 2.5 3.0 4.0 6.0 10.0 Tailing

4.36 68.26 4.19 78.15 3.88 73.35 3.74 72.86 3.95 68.51 4.80 71.85 5.21 68.93 5.75 65.80 10.11 33.29

12.84 11.89 11.40 12.01 12.00 13.97 13.87 13.80 11.08

4.71 6.45 7.22 8.90 8.97 9.25 9.57 9.53 11.65

72.85 71.70 58.34 46.09 37.07 34.91 33.43 30.48 16.38

9.65 9.65 8.97 9.31 9.16 8.32 9.14 7.86 8.12

4.67 67.22 5.44 71.52 6.16 65.46 6.45 53.67 6.49 52.06 7.06 44.41 7.54 43.50 7.89 37.55 10.41 15.54

9.58 8.36 9.56 9.36 7.26 6.90 7.86 7.26 8.31

4.73 72.99 5.77 70.48 6.86 57.51 7.41 49.69 7.76 45.67 7.68 40.92 7.60 38.36 7.58 36.18 10.16 22.94

7.99 7.61 8.39 7.79 8.65 7.14 8.66 8.55 7.54

decreasing mineral particle size. At 10 min, recovery rates of 25, 53, 55 and 59% are achieved for the four sizes tested. As expected, when oleic acid concentration is increased to 2000 g/t (Fig. 9), the results are slightly higher than those obtained at 1000 g/t, but follow the same pattern with regard to variation with particle size as stated above. From the standpoint of P 2O5, CaCO3 and SiO2 content, the results for 1000 g/t (Table 2) clearly show that calcite is the species that floats best given that the concentration found in the concentrate is always higher than in the tailing, and for particle sizes between 160 and 200 mesh, concentrate remain at 65–78% compared with only 33.29% in the tailing. P 2 O5 content in the various concentrates ranges between 3.74 and 5.75, while in the tailing it is 10.11%; SiO 2 content, although this does not alter to any considerable degree, is slightly higher in the concentrate (11.40–13.97 compared with 11.08%). For 200/230 mesh and lower particle sizes, SiO 2 content remains within a similar range as described above. However, CaCO 3 content in the various concentrates and the tailing are highly indicative of how the process occurs, i.e. during the first stage of flotation (1 min) calcite content in the concentrate is much higher, up to over 72% in some cases, while at 10 min flotation times, the content of the same mineral drops to below 31%. P2O5 content, although on a lesser scale with variations between 4 and 10%, is low during the first stages of flotation and rises towards the end. The content in the tailing is always around 10%. Similar results are achieved when oleic acid concentrations are raised to a 2000 g/t. 4. CONCLUSIONS

The results obtained confirm that for particle sizes of 200/230 mesh and under, the carbonate separates during the first moments of flotation (1.5 min) and can reach up

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to 80% content, while these content levels drop towards the end of flotation (10 min) down to around 31% at the least. The opposite occurs with P 2O5, at the beginning of the operation — content percentages are low (4–6%), and then rise during the operation for 200/230 mesh and under to reach 8–9% at the end of flotation. The above seems to suggest that very short flotation times (less than 2 min) should be used when separating CaCO 3 and P2O5 through flotation, given that it is in this time range that the greatest differences are achieved.

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Flotation of Low-grade Phosphate Ore

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