Beneficiation of a Sedimentary Phosphate Ore by a Combination of Spiral Gravity and Direct-Reverse Flotation

Xin Liu 1, Yimin Zhang 1,2,3,*, Tao Liu 1,2,3, Zhenlei Cai 1,2,3, Tiejun Chen 1,2,3 and Kun Sun 1 1 College of Resource and Environmental Engineering, Wuhan University of Science and Technology, Wuhan 430081, China; xin_liu1314@163.com (X.L.); tkliutao@126.com (T.L.); ande559@163.com (Z.C.); ctj_56@163.com (T.C.); sunkuniq@163.com (K.S.) 2 Hubei Collaborative Innovation Center for High Efficient Utilization of Vanadium Resources, Wuhan 430081, China 3 Hubei Provincial Engineering Technology Research Center of High efficient Cleaning Utilization for Shale Vanadium Resource, Wuhan 430081, China * Correspondence: zym126135@126.com; Tel./Fax: +86-27-6886-2057


Introduction
Phosphate rocks are vital non-renewable resources and are essential components in fertilizer production.There is increasing demand for phosphate rock for fertilizers, animal feed stocks, food-grade phosphates, and other industrial uses [1][2][3].About 95% of the phosphate produced around the world is consumed in the fertilizer industry [4,5].China has about 20 billion tonnes of abundant phosphate ore reserves (including reserve base) ranking it second in the world [6].The phosphate ores in China can be classified as igneous rock, metasedimentary, and sedimentary, and each are significantly different in their flotation characteristics.The igneous apatite ores and metasedimentary phosphate rocks have relatively good floatability.However, these reserves are rapidly being depleted, prompting the development of recovery methods for the more abundant sedimentary ores [4,7].The sedimentary ores contain apatite with a very small grain size, typically micro-to optically-visible, which can be highly intergrown with gangue materials and are, thus, called cryptocrystal, or classified as "colloidal phosphate ore" (collophanite) in China.This implies that the sedimentary ores are difficult to float and require fine grinding.
Phosphate ore processing techniques are dependent on the type of phosphate minerals and associate gangue.To meet the requirement of fertilizer industry, phosphate ore should be enriched to near 30% P 2 O 5 [8].Flotation has been used in industry for more than half century as the primary technique for upgrading phosphate.More than 60% of the commercial phosphate in the world is produced by flotation [9].The flotation process not only upgrading the P 2 O 5 grade but also reducing the particle size to 100 mesh content more than 90% [10].The present techniques used in flotation of the sedimentary phosphate ore are generally classified into three categories: (i) direct-reverse flotation (or reverse-direct flotation); (ii) single-reverse flotation, and (iii) double-reverse flotation [2,11,12].The direct-reverse method is more attractive, as it has been proved to be applicable to most phosphate ores in China [7].In this process, the siliceous minerals are discharged by direct flotation while the calcareous minerals are discharged by reverse flotation.Although the direct-reverse process has been improved with the use of more efficient collectors, reagent consumption has also increased significantly.Therefore, there is an imperative to improve this process or develop alternate processing methods.
Gravity separation is an important unit process due to its advantages of easy manipulation and low cost.Some research on the application of gravity separation for sedimentary phosphate ore has been conducted.Yang [13] investigated a gravity-flotation process.The results showed that shaking table gravity separation could reduce ore mass flow to the flotation process.However, this combined process had high capital and operating costs.Attrition scrubbing techniques have been successfully used to upgrade some mid-grade, weathered, sedimentary phosphate ores in the Yunnan province, China [14].This technique was used when the main gangue minerals were clays.However, with a decline of the ore grade and an increase of the content of impurities, attrition scrubbing processes have gradually lost their place.Another successfully commercialized gravity technique is the heavy medium cyclone process, used for discharging coarse dolomite at Yihua Group, Hubei province, China [14,15].The gravity concentrate was then subjected to double-reverse flotation for further discharging of dolomite.However, it has been shown to not be suitable for processing fine particles.
Among various kinds of gravity separators, the spiral concentrator is considered to be one of the most efficient and simple operation units.Since their fully commercial use in the early 1940s, spirals have proved to be a cost-effective and efficient method for concentrating a variety of ores [16,17].In recent years, fine mineral spiral separators have been successfully used for processing ever-finer materials [18][19][20].Due to the low settling velocities of very fine mineral particles, smaller bed depths and laminar flows are required for efficient separation [18].Accordingly, a fine mineral spiral separator has a significantly smaller ratio of pitch to diameter (P/D) than the traditional spirals [21].However, there is little information on recovery of apatite or collophanite from ultra-fine phosphate ore by fine spiral separators.
The objective of this work was to investigate a simple gravity process for the possible pre-recovery of collophanite from direct flotation allowing a reduction in reagent consumptions.The gravity concentrator used was a fine mineral spiral, with a P/D ratio of only 0.36.The study focused on the reagent consumptions of two different flowsheets, the conventional direct-reverse flotation process and the gravity-flotation combined process.A new gravity-flotation combined process was proposed and tested, which significantly reduced the cost for producing 1 t of concentrate.

Materials
The sedimentary phosphate ore sample used in this work was collected from Yihua Group Co., Ltd., Yichang, China.The results of chemical composition analysis and mineral composition are shown in Tables 1 and 2. Around 200 kg of representative ore samples was crushed to below 3 mm with a two-stage jaw crusher and a one-stage roll crusher.The materials were then well mixed and divided into 1 kg samples for separation studies.The chemical analysis was performed with the Xios advanced X-ray fluorescence (XRF) analyzer, the results are tabulated in Table 1.Optical microscopy and X-ray diffraction (XRD) studies were performed to define the main minerals and their interlocking characteristics.The results of optical mineralogy are showed in Figure 1.Collophanite grains are polygonal, plate-like, and are usually between 0.05 and 0.8 mm.Small apatite grains (0.01-0.05 mm) could be found in some collophanite grains (Figure 1b).Some of the gangue minerals are occurs as endogangue in the collophanite grains as shown in Figure 1a,c,d.The majority of gangue minerals are appear between the collophanite grains.These gangue minerals are clay minerals, the aggregate of quartz, feldspar, and dolomite.The grains of quartz and feldspar are small and difficult to distinguish between them.Their aggregate is between 0.03 and 0.1 mm, and is usually in the exogangue (Figure 1d) and sporadically distributed in the collophanite grains (Figure 1c).It can be seen from Figure 1c that the dolomite minerals are hypidiomorphic granular structure, scattered, and grains size are usually between 0.03 and 0.1 mm.Dolomite as exogangue (Figure 1c) and endogangue (Figure 1d) exist between and in the collophanite grains.Clay mineral pellets are mainly composed of illite, kaolinite.They usually appear in the form of crystal stock, either or as endogangue in collophanite grains (Figure 1a).In addition to this, there are slight iron minerals like pyrite (Figure 1e), magnetite (Figure 1f) and hematite created small grains (0.01-0.03 mm) which sporadically distribute in collophanite grains, quartz and feldspar aggregates, and clay minerals aggregates.Optical microscopy and X-ray diffraction (XRD) studies were performed to define the main minerals and their interlocking characteristics.The results of optical mineralogy are showed in Figure 1.Collophanite grains are polygonal, plate-like, and are usually between 0.05 and 0.8 mm.Small apatite grains (0.01-0.05 mm) could be found in some collophanite grains (Figure 1b).Some of the gangue minerals are occurs as endogangue in the collophanite grains as shown in Figure 1a,c,d.The majority of gangue minerals are appear between the collophanite grains.These gangue minerals are clay minerals, the aggregate of quartz, feldspar, and dolomite.The grains of quartz and feldspar are small and difficult to distinguish between them.Their aggregate is between 0.03 and 0.1 mm, and is usually in the exogangue (Figure 1d) and sporadically distributed in the collophanite grains (Figure 1c).It can be seen from Figure 1c that the dolomite minerals are hypidiomorphic granular structure, scattered, and grains size are usually between 0.03 and 0.1 mm.Dolomite as exogangue (Figure 1c) and endogangue (Figure 1d) exist between and in the collophanite grains.Clay mineral pellets are mainly composed of illite, kaolinite.They usually appear in the form of crystal stock, either or as endogangue in collophanite grains (Figure 1a).In addition to this, there are slight iron minerals like pyrite (Figure 1e), magnetite (Figure 1f) and hematite created small grains (0.01-0.03 mm) which sporadically distribute in collophanite grains, quartz and feldspar aggregates, and clay minerals aggregates.The results of optical mineralogy and the XRD analysis (Figure 2) revealed the major phases of minerals which to be apatite, quartz, dolomite, clay minerals, and pyrite.The mineral composition of the sample was obtained by the comprehensive analysis of the chemical composition, optical microscopy and XRD.The results are shown in Table 2.The density of the main minerals is also tabulated in Table 2 [22].The samples were firstly wet ground in a HLXMQ-Φ240 × 90 laboratory ball mill from Wuhan Exploring Machinery Factory (Wuhan, China) at 50 wt % solids, until a particle size distribution of 70% passing 74 μm was achieved.The ground product was then subjected to characterization studies and separation studies.The particle size fraction analysis which was carried out using a Mastersizer 2000 laser particle characterization system (Malvern Instruments Ltd., Malvern, UK), showing that the specific surface of the ground product was 0.78 m 2 /g, and the d10, d50, and d90 (diameter at the cumulative undersize of 10%, 50%, and 90%) was 2, 51 and 115 μm, respectively.The liberation of collophanite was analyzed.To start with, a representative sample of particles weighing approximately 5 g was taken from +30 μm size fractions of the ground product.The −30 μm latter size fraction was not prepared for liberation analysis in a reflecting microscope, due to magnification restrictions, which makes it difficult to definitely distinguish mineral particles.The results showed that the liberation degree in +30 μm is 70.4%.The viscosity of slurry was measured by NDJ-5S viscometer (Hengping Scientific Instruments Ltd., Shanghai, China).The results showed that, while the solid content between 10% and 35%, the viscosity increased slowly from 1.168 to 1.176 mPa•s as the solid content increased.Surface tension measurement was conducted by the Du Noüy ring method.The results showed that the surface tension of the slurry was 63 mN/m at reverse flotation conditions, while 54 mN/m at direct flotation conditions.The results of optical mineralogy and the XRD analysis (Figure 2) revealed the major phases of minerals which to be apatite, quartz, dolomite, clay minerals, and pyrite.The mineral composition of the sample was obtained by the comprehensive analysis of the chemical composition, optical microscopy and XRD.The results are shown in Table 2.The density of the main minerals is also tabulated in Table 2 [22].The results of optical mineralogy and the XRD analysis (Figure 2) revealed the major phases of minerals which to be apatite, quartz, dolomite, clay minerals, and pyrite.The mineral composition of the sample was obtained by the comprehensive analysis of the chemical composition, optical microscopy and XRD.The results are shown in Table 2.The density of the main minerals is also tabulated in Table 2 [22].The samples were firstly wet ground in a HLXMQ-Φ240 × 90 laboratory ball mill from Wuhan Exploring Machinery Factory (Wuhan, China) at 50 wt % solids, until a particle size distribution of 70% passing 74 μm was achieved.The ground product was then subjected to characterization studies and separation studies.The particle size fraction analysis which was carried out using a Mastersizer 2000 laser particle characterization system (Malvern Instruments Ltd., Malvern, UK), showing that the specific surface of the ground product was 0.78 m 2 /g, and the d10, d50, and d90 (diameter at the cumulative undersize of 10%, 50%, and 90%) was 2, 51 and 115 μm, respectively.The liberation of collophanite was analyzed.To start with, a representative sample of particles weighing approximately 5 g was taken from +30 μm size fractions of the ground product.The −30 μm latter size fraction was not prepared for liberation analysis in a reflecting microscope, due to magnification restrictions, which makes it difficult to definitely distinguish mineral particles.The results showed that the liberation degree in +30 μm is 70.4%.The viscosity of slurry was measured by NDJ-5S viscometer (Hengping Scientific Instruments Ltd., Shanghai, China).The results showed that, while the solid content between 10% and 35%, the viscosity increased slowly from 1.168 to 1.176 mPa•s as the solid content increased.Surface tension measurement was conducted by the Du Noüy ring method.The results showed that the surface tension of the slurry was 63 mN/m at reverse flotation conditions, while 54 mN/m at direct flotation conditions.The samples were firstly wet ground in a HLXMQ-Φ240 ˆ90 laboratory ball mill from Wuhan Exploring Machinery Factory (Wuhan, China) at 50 wt % solids, until a particle size distribution of 70% passing 74 µm was achieved.The ground product was then subjected to characterization studies and separation studies.The particle size fraction analysis which was carried out using a Mastersizer 2000 laser particle characterization system (Malvern Instruments Ltd., Malvern, UK), showing that the specific surface of the ground product was 0.78 m 2 /g, and the d10, d50, and d90 (diameter at the cumulative undersize of 10%, 50%, and 90%) was 2, 51 and 115 µm, respectively.The liberation of collophanite was analyzed.To start with, a representative sample of particles weighing approximately 5 g was taken from +30 µm size fractions of the ground product.The ´30 µm latter size fraction was not prepared for liberation analysis in a reflecting microscope, due to magnification restrictions, which makes it difficult to definitely distinguish mineral particles.The results showed that the liberation degree in +30 µm is 70.4%.The viscosity of slurry was measured by NDJ-5S viscometer (Hengping Scientific Instruments Ltd., Shanghai, China).The results showed that, while the solid content between 10% and 35%, the viscosity increased slowly from 1.168 to 1.176 mPa¨s as the solid content increased.Surface tension measurement was conducted by the Du Noüy ring method.The results showed that the surface tension of the slurry was 63 mN/m at reverse flotation conditions, while 54 mN/m at direct flotation conditions.Commercial water glass (WG) from Xiangxing Co. Ltd. (Loudi, China) of modulus 3.1 was used as the depressant for silicate minerals and dispersant, which had a purity of 97%.Na 2 CO 3 from Xilong Chemical Co., Ltd.(Wuhan, China) with purity of 99% was used as a pH modifier in direct flotation.Sulfuric acid (SA) from Pingmei Group Co. Ltd. (Pingdingshan, China) was used as a pH modifier (98% purity) in reverse flotation.Commercial product CXY-P (a fatty acid mixture) from Chuxiang Phosphorus Technology Ltd. (Wuhan, China) was used as a collector.
The water used was the potable water in Wuhan, China.

Methods
The process with the gravity and the flotation was termed as the "new process".The diagram of the actual process which was applied during 2014 is shown in Figure 3a.The diagram of the conventional direct-reverse flotation configuration is shown in Figure 3b.Spiral separation was used to improve the economics of the process and the new combined process is illustrated in Figure 3c.The water used was the potable water in Wuhan, China.

Methods
The process with the gravity and the flotation was termed as the "new process".The diagram of the actual process which was applied during 2014 is shown in Figure 3a.The diagram of the conventional direct-reverse flotation configuration is shown in Figure 3b.Spiral separation was used to improve the economics of the process and the new combined process is illustrated in Figure 3c.

Gravity Separation Tests
The gravity tests are comprised classification and spiral separation.In order to avoid the negative effects of coarse particles on the spiral separation, classification was firstly used for recovery of +150 μm size fraction particles from the feed.Then spiral separation was carried out.The separator used for the tests was a five-turn laboratory spiral.The design data of the spiral are given in Table 3 and the set-up used for the experiments is shown in Figure 4.A manual valve combined with a rotameter was used to adjust the flow of wash water for the spiral.The wash water was only supplied when sampling.Batch experiments were conducted to investigate different influences; solid content, feed rate, and wash water.

Gravity Separation Tests
The gravity tests are comprised classification and spiral separation.In order to avoid the negative effects of coarse particles on the spiral separation, classification was firstly used for recovery of +150 µm size fraction particles from the feed.Then spiral separation was carried out.The separator used for the tests was a five-turn laboratory spiral.The design data of the spiral are given in Table 3 and the set-up used for the experiments is shown in Figure 4.A manual valve combined with a rotameter was used to adjust the flow of wash water for the spiral.The wash water was only supplied when sampling.Batch experiments were conducted to investigate different influences; solid content, feed rate, and wash water.

Flotation Tests
Flotation tests were carried out using a XFDIV (equipment model) single flotation machine from Jilin Provincial Machine Factory of Ore Exploration (Changchun, China).The impeller speed was 1750 rpm and flotation was conducted at a pulp temperature of 25 °C.A 1.0 L flotation cell was prepared for the direct flotation tests, and a 0.75 L flotation cell for the reverse flotation.For each direct flotation test, the pulp containing 330 g of feed ore was transferred into the flotation cell and stirred for 1 min.Sodium carbonate was added into the pulp to achieve a pH at 10 with 2 min of conditioning time.Then WG was added, and conditioned for another 2 min.Next, the collector CXY-P was added and the slurry was stirred for 2 min.All conditioning processes were conducted in the absence of airflow.For each reverse flotation test, the feed was the concentrate from ether the gravity or the direct flotation.SA was firstly added to achieve a pH at 4.5 with 0.5 min conditioning time.CXY-P was added and conditioned for 2 min before flotation.Batch experiments were conducted to find the optimal dosages of reagents by single factor test method, and using the parallel test to increase the stability.The optimal flotation conditions are shown in Figure 5.

Flotation Tests
Flotation tests were carried out using a XFDIV (equipment model) single flotation machine from Jilin Provincial Machine Factory of Ore Exploration (Changchun, China).The impeller speed was 1750 rpm and flotation was conducted at a pulp temperature of 25 ˝C.A 1.0 L flotation cell was prepared for the direct flotation tests, and a 0.75 L flotation cell for the reverse flotation.For each direct flotation test, the pulp containing 330 g of feed ore was transferred into the flotation cell and stirred for 1 min.Sodium carbonate was added into the pulp to achieve a pH at 10 with 2 min of conditioning time.Then WG was added, and conditioned for another 2 min.Next, the collector CXY-P was added and the slurry was stirred for 2 min.All conditioning processes were conducted in the absence of airflow.For each reverse flotation test, the feed was the concentrate from ether the gravity or the direct flotation.SA was firstly added to achieve a pH at 4.5 with 0.5 min conditioning time.CXY-P was added and conditioned for 2 min before flotation.Batch experiments were conducted to find the optimal dosages of reagents by single factor test method, and using the parallel test to increase the stability.The optimal flotation conditions are shown in Figure 5.

Flotation Tests
Flotation tests were carried out using a XFDIV (equipment model) single flotation machine from Jilin Provincial Machine Factory of Ore Exploration (Changchun, China).The impeller speed was 1750 rpm and flotation was conducted at a pulp temperature of 25 °C.A 1.0 L flotation cell was prepared for the direct flotation tests, and a 0.75 L flotation cell for the reverse flotation.For each direct flotation test, the pulp containing 330 g of feed ore was transferred into the flotation cell and stirred for 1 min.Sodium carbonate was added into the pulp to achieve a pH at 10 with 2 min of conditioning time.Then WG was added, and conditioned for another 2 min.Next, the collector CXY-P was added and the slurry was stirred for 2 min.All conditioning processes were conducted in the absence of airflow.For each reverse flotation test, the feed was the concentrate from ether the gravity or the direct flotation.SA was firstly added to achieve a pH at 4.5 with 0.5 min conditioning time.CXY-P was added and conditioned for 2 min before flotation.Batch experiments were conducted to find the optimal dosages of reagents by single factor test method, and using the parallel test to increase the stability.The optimal flotation conditions are shown in Figure 5.

Gravity Separation Results
The results of particle size analyses of the feed (Table 4) showed that the P 2 O 5 grade of +150 µm in the feed was more than 26%.It could be collected as a separate concentrate by classification.In the spiral separation process, the spiral feed (´150 µm of the feed) was separated into seven streams, numbered from 1 to 7. The width of every stream was 24 mm.A cross-section of the spiral trough flow is shown in Figure 6 and illustrates the separation process.Optimized test conditions showed a pulp density of 18 wt % solids, a wash water flow rate of 30 L/h, and a feed rate of 150 L/h.

Gravity Separation Results
The results of particle size analyses of the feed (Table 4) showed that the P2O5 grade of +150 μm in the feed was more than 26%.It could be collected as a separate concentrate by classification.In the spiral separation process, the spiral feed (−150 μm of the feed) was separated into seven streams, numbered from 1 to 7. The width of every stream was 24 mm.A cross-section of the spiral trough flow is shown in Figure 6 and illustrates the separation process.Optimized test conditions showed a pulp density of 18 wt % solids, a wash water flow rate of 30 L/h, and a feed rate of 150 L/h.  Figure 7 shows that as the product number increased, P2O5 content decreased sharply down to number 4, then it slowly increased above that.Meanwhile, SiO2, Al2O3, and MgO contents were concentrated in light products (high product number).Figure 7 shows that as the product number increased, P 2 O 5 content decreased sharply down to number 4, then it slowly increased above that.Meanwhile, SiO 2 , Al 2 O 3 , and MgO contents were concentrated in light products (high product number).

Gravity Separation Results
The results of particle size analyses of the feed (Table 4) showed that the P2O5 grade of +150 μm in the feed was more than 26%.It could be collected as a separate concentrate by classification.In the spiral separation process, the spiral feed (−150 μm of the feed) was separated into seven streams, numbered from 1 to 7. The width of every stream was 24 mm.A cross-section of the spiral trough flow is shown in Figure 6 and illustrates the separation process.Optimized test conditions showed a pulp density of 18 wt % solids, a wash water flow rate of 30 L/h, and a feed rate of 150 L/h.  Figure 7 shows that as the product number increased, P2O5 content decreased sharply down to number 4, then it slowly increased above that.Meanwhile, SiO2, Al2O3, and MgO contents were concentrated in light products (high product number).Figure 8 illustrates the P 2 O 5 cumulative grade vs. product number curve.When the product number increased above 2, the P 2 O 5 cumulative grade decreased significantly.This data suggest that the concentrate splitter position at product number 2 could achieve higher selectivity between collophanite and gangue minerals.
Figure 8 illustrates the P2O5 cumulative grade vs. product number curve.When the product number increased above 2, the P2O5 cumulative grade decreased significantly.This data suggest that the concentrate splitter position at product number 2 could achieve higher selectivity between collophanite and gangue minerals.The separation results (Table 5) also indicated that the collophanite particles were concentrated in the high density products (heavies) and gangue particles separated to low density streams.It can be observed that the heavies of the spirals were enriched to 28.22% P2O5 with a distribution of 52.8% in 40.3% of the weight.

Flotation Tests Results
Table 6 shows the results of the flotation tests.A concentrate of 30.56%P2O5 was produced at a P2O5 recovery of 86.3% by the new process.With a given P2O5 recovery rate (86%), the new process produced a concentrate with a slight little lower P2O5 grade (30.56%) than the conventional process (31.69%).The separation results (Table 5) also indicated that the collophanite particles were concentrated in the high density products (heavies) and gangue particles separated to low density streams.It can be observed that the heavies of the spirals were enriched to 28.22% P 2 O 5 with a distribution of 52.8% in 40.3% of the weight.

Flotation Tests Results
Table 6 shows the results of the flotation tests.A concentrate of 30.56%P 2 O 5 was produced at a P 2 O 5 recovery of 86.3% by the new process.With a given P 2 O 5 recovery rate (86%), the new process produced a concentrate with a slight little lower P 2 O 5 grade (30.56%) than the conventional process (31.69%).

Flowsheet Test Results
The flowsheet for combined gravity-flotation (Figure 9) shows that the P 2 O 5 grade of the final concentrate was 30.41%, while the P 2 O 5 recovery was 91.5%.In this process, 1.8% and 51.0% collophanite were pre-collected by gravity separation using classification and spiral separators, respectively.The lights from spiral separation was then subjected to direct flotation.The mixed concentrate from the gravity-direct flotation was subjected to reverse flotation.Combined with the XRE analysis (Figure 2) and the XRF analysis (Figure 10) of concentrates and raw ore, it is showed that Al primarily originated from clay mineral.Compared to the conventional process, the new process produced a concentrate with a slight little lower Al 2 O 3 content.It is indicated that the new process can improve the efficiency of discarding clay minerals.

Flowsheet Test Results
The flowsheet for combined gravity-flotation (Figure 9) shows that the P2O5 grade of the final concentrate was 30.41%, while the P2O5 recovery was 91.5%.In this process, 1.8% and 51.0% collophanite were pre-collected by gravity separation using classification and spiral separators, respectively.The lights from spiral separation was then subjected to direct flotation.The mixed concentrate from the gravity-direct flotation was subjected to reverse flotation.Combined with the XRE analysis (Figure 2) and the XRF analysis (Figure 10) of concentrates and raw ore, it is showed that Al primarily originated from clay mineral.Compared to the conventional process, the new process produced a concentrate with a slight little lower Al2O3 content.It is indicated that the new process can improve the efficiency of discarding clay minerals.In the actual process, which was applied during 2014, beneficiation recovery is about 69% at reagent costs of 14.6 CNY/1 t of concentrate and raw ore cost of 209 CNY.A comparison of the costs in the actual process, the conventional process and the new process was made.The results are shown in Figure 11.Compared to the actual process, in the new process, the raw ore cost for producing 1 t of concentrate decreases by 38.8 CNY.However, the reagent costs only increase 7.5 CNY/1 t of concentrate.The net benefit is 31.3CNY/1 t of concentrate and boosts the profits by 39%, whilst the conventional process costs more than the actual process.

Conclusions
1.The results of this study showed that the fine spiral of a 0.36 P/D ratio was an useful separator for pre-recovery of collophanite from Yichang ultra-fine sedimentary phosphate ore.A final concentrate with a grade of 30.41%P2O5 and a recovery of 91.5% can be produced by a combined process of gravity and direct-reverse flotation.2. Compared to the actual process applied during the year 2014, the use of gravity concentrator in the flowsheet achieved significant cost savings.The net benefit is 31.3CNY/1 t concentrate and boosts the profits by 39%, whilst the conventional process costs 10.8 CNY more than the actual process for producing 1 t of concentrate.In the actual process, which was applied during 2014, beneficiation recovery is about 69% at reagent costs of 14.6 CNY/1 t of concentrate and raw ore cost of 209 CNY.A comparison of the costs in the actual process, the conventional process and the new process was made.The results are shown in Figure 11.Compared to the actual process, in the new process, the raw ore cost for producing 1 t of concentrate decreases by 38.8 CNY.However, the reagent costs only increase 7.5 CNY/1 t of concentrate.The net benefit is 31.3CNY/1 t of concentrate and boosts the profits by 39%, whilst the conventional process costs more than the actual process.In the actual process, which was applied during 2014, beneficiation recovery is about 69% at reagent costs of 14.6 CNY/1 t of concentrate and raw ore cost of 209 CNY.A comparison of the costs in the actual process, the conventional process and the new process was made.The results are shown in Figure 11.Compared to the actual process, in the new process, the raw ore cost for producing 1 t of concentrate decreases by 38.8 CNY.However, the reagent costs only increase 7.5 CNY/1 t of concentrate.The net benefit is 31.3CNY/1 t of concentrate and boosts the profits by 39%, whilst the conventional process costs more than the actual process.

Conclusions
1.The results of this study showed that the fine spiral of a 0.36 P/D ratio was an useful separator for pre-recovery of collophanite from Yichang ultra-fine sedimentary phosphate ore.A final concentrate with a grade of 30.41%P2O5 and a recovery of 91.5% can be produced by a combined process of gravity and direct-reverse flotation.2. Compared to the actual process applied during the year 2014, the use of gravity concentrator in the flowsheet achieved significant cost savings.The net benefit is 31.3CNY/1 t concentrate and boosts the profits by 39%, whilst the conventional process costs 10.8 CNY more than the actual process for producing 1 t of concentrate.

1.
The results of this study showed that the fine spiral of a 0.36 P/D ratio was an useful separator for pre-recovery of collophanite from Yichang ultra-fine sedimentary phosphate ore.A final concentrate with a grade of 30.41%P 2 O 5 and a recovery of 91.5% can be produced by a combined process of gravity and direct-reverse flotation.

2.
Compared to the actual process applied during the year 2014, the use of gravity concentrator in the flowsheet achieved significant cost savings.The net benefit is 31.3CNY/1 t concentrate and boosts the profits by 39%, whilst the conventional process costs 10.8 CNY more than the actual process for producing 1 t of concentrate.
glass (WG) from Xiangxing Co. Ltd. (Loudi, China) of modulus 3.1 was used as the depressant for silicate minerals and dispersant, which had a purity of 97%.Na2CO3 from Xilong Chemical Co., Ltd.(Wuhan, China) with purity of 99% was used as a pH modifier in direct flotation.Sulfuric acid (SA) from Pingmei Group Co. Ltd. (Pingdingshan, China) was used as a pH modifier (98% purity) in reverse flotation.Commercial product CXY-P (a fatty acid mixture) from Chuxiang Phosphorus Technology Ltd. (Wuhan, China) was used as a collector.

Figure 3 .
Figure 3. (a) Diagram of the actual process (2014); (b) the conventional process and (c) the new gravity and flotation combined process.

Figure 3 .
Figure 3. (a) Diagram of the actual process (2014); (b) the conventional process and (c) the new gravity and flotation combined process.

Figure 5 .
Figure 5.The flotation flowsheet for the conventional process (a) and the new process (b).

Figure 5 .
Figure 5.The flotation flowsheet for the conventional process (a) and the new process (b).Figure 5.The flotation flowsheet for the conventional process (a) and the new process (b).

Figure 5 .
Figure 5.The flotation flowsheet for the conventional process (a) and the new process (b).Figure 5.The flotation flowsheet for the conventional process (a) and the new process (b).

Figure 6 .
Figure 6.A sectional view of the spiral trough flow.

Figure 7 .
Figure 7.Chemical compositions of the products from spiral separation.

Figure 6 .
Figure 6.A sectional view of the spiral trough flow.

Figure 6 .
Figure 6.A sectional view of the spiral trough flow.

Figure 7 .
Figure 7.Chemical compositions of the products from spiral separation.Figure 7. Chemical compositions of the products from spiral separation.

Figure 7 .
Figure 7.Chemical compositions of the products from spiral separation.Figure 7. Chemical compositions of the products from spiral separation.

Figure 10 .
Figure 10.Chemical compositions of concentrates and raw ore.

Figure 11 .
Figure 11.Benefit of the new process and the conventional process, compared to the actual process (2014).

Figure 10 .
Figure 10.Chemical compositions of concentrates and raw ore.

Figure 10 .
Figure 10.Chemical compositions of concentrates and raw ore.

Figure 11 .
Figure 11.Benefit of the new process and the conventional process, compared to the actual process (2014).

Figure 11 .
Figure 11.Benefit of the new process and the conventional process, compared to the actual process (2014).

Table 1 .
Chemical composition of raw ore.

Table 2 .
Mineral composition of raw ore.

Table 1 .
Chemical composition of raw ore.

Table 2 .
Mineral composition of raw ore.

Table 3 .
Physical structure parameters of the spiral.

Table 3 .
Physical structure parameters of the spiral.

Table 4 .
Particle size and chemical analyses of the feed (Dist.: distribution).

Table 4 .
Particle size and chemical analyses of the feed (Dist.: distribution).

Table 4 .
Particle size and chemical analyses of the feed (Dist.: distribution).

Table 6 .
Open circuit flotation results for the conventional process and the new process.

Table 6 .
Open circuit flotation results for the conventional process and the new process.