Conventional froth flotation is one of the most widely used mineral beneficiation processes. Unfortunately, long residence times are often required to achieve complete flotation. These slow flotation kinetics have been associated with the hydrodynamics within the flotation cell. While flotation cell manufacturers have optimized the mixing hydrodynamics, little effort has been made to improve the flotation hydrodynamics within the flotation cell.

A conventional flotation cell has both a turbulent region, where bubble-particle attachment occurs, and a quiescent region, where the ore-laden froth is removed from the cell. The flotation rate of a single particle in this environment is a function of the probabilities of bubble-particle collision, attachment, and the motion of the bubble moving into the froth. To increase the flotation rate of a flotation cell, the number of bubble-particle collisions must be increased. One way to increase the number of collisions is to increase the number of bubbles, either by decreasing the bubble size or by increasing the air flow rate to the flotation cell. Unfortunately, the bubble size is limited by the mechanical action in a conventional cell, and at high air flow rates, the mechanical cell’s agitation is less effective at breaking up the air into small bubbles. Another way to increase the number of collisions is to raise the turbulence within the cell by increasing the fluid velocity. Increasing the fluid velocity increases the relative velocity of the bubbles to the particles and improves the probability of bubble-particle collision. However, increasing the turbulence can also cause entrainment, where the bubble-particle aggregate is removed from the flotation call along with the tailings. Increasing the turbulence can also cause detachment, where the flow conditions dislodge the particles from the bubbles. Thus, the process of attaching the particle to the bubble and the subsequent separation of the bubble from the pulp interact significantly in a conventional flotation cell, making it virtually impossible to optimize one of the processes without sacrificing the effectiveness of the other.

As a part of the Bureau of Mines’ efforts to improve the efficiency of the domestic minerals industry, the Bureau has developed a rapid froth flotation system which separates flotation into two discrete unit operations: bubble-particle attachment and bubble-pulp separation (figure 1). In the bubble-particle attachment unit, a bubble slurry is mixed with an ore slurry under highly turbulent conditions. The mixture is then rapidly fed into a quiescent bubble-pulp separation unit where the mineral laden bubbles are quickly separated from the pulp.

Experimental Procedure

In this investigation, an in-line static mixer was evaluated as a bubble-particle attachment unit. The bubble slurry was pumped from the bubble generator. In the in-line static mixer, the bubbles and particles passed through each other at high velocities, causing rapid bubble-particle contact. The froth-pulp mixture immediately entered the shallow-depth froth separator (figure 2) through a sparger at the top of the unit. The sparger simply dispersed the froth-pulp mixture which provided the quiescent conditions necessary for adequate bubble-pulp separation. The relatively large diameter froth separator had a nominal depth of flow (approx. 2 cm) which minimized the rising time of even the smallest bubbles (100 µm) . As the mineral pulp moved outward from the sparger, the ore-laden bubbles quickly rose to the top of the pulp layer and overflowed at the edge of the separator. The depth of this horizontal flow was fairly shallow, because the sparger distributes the pulp at the top of the separator. The effective volume of the froth separator was the disk shaped region at the top and was roughly equal to the surface area times the depth of the flow. Bubbles must rise out of the pulp during the time that the pulp is in this effective region. This time is referred to as the effective bubble residence time, because any bubbles that have not risen to the surface by the time the pulp reaches the outer edge of the froth separator will be swept from the separator with the tailings as the pulp exits the unit.

The effective bubble residence time may be calculated using the following equation:

where df is the depth of the flow, As is the surface area of the froth separator, and Q is the volume flow rate. The effective volume largely depends upon the type of sparger used to distribute the flow radially. By using the same sparger throughout the test work, the effective depth remained relatively constant even as the flow rate varied. Therefore, for this research, the same sparger was used at every flow rate and with every froth separator. This kept the df relatively constant as the pulp radiated fromt he sparger. The unfloated pulp moved out the bottom of the unit.

To evaluate the effectiveness of the rapid flotation system, a western porphyry copper and a phosphate-bearing amine flotation tailings ore were tested. Conventional flotation tests were also carried out in a Denver flotation cell in order to compare the results with those obtained from the rapid flotation technique. Tests were conducted at various impeller speeds and air flow rates to obtain to optimum conditions for flotation. Best results were obtained when the impeller cell was set at 1,100 r/min and air was introduced at 4 L/min. The same reagent scheme and pH conditions were used in the rapid and conventional systems. Timed samples of the flotation concentrate were taken to determine the flotation kinetics of the copper ore and the amine tailings sample in the Denver flotation cell.

The rapid flotation system was designed to agitate the bubbles and ore pulp as they flowed through an in-line static mixer, and to continuously remove the mineral laden bubbles from the ore pulp. The flotation system consisted of a slurry reservoir, a bubble generator, a bubble-particle attachment unit and a froth separator. The ore was conditioned with reagents in the slurry reservoir and pumped through the in-line static mixer along with the bubble slurry. The bubble pulp mixture, from the in-line static mixer, was then sent to the froth separator where the concentrate and tailings products were separated. The flotation process was continued until the ore slurry in the reservoir was depleted. At that time, 2 L of wash water was added to the ore reservoir to flush the system of any remaining ore. The concentrate and tailings were dried, weighed, and analyzed.

The bubble generator was a conventional flotation cell (10 L) with the normal impeller replaced by a 16.5-cm-diam high-speed spinning disk, rotating at 3,300 r/min. Air striking the spinning disk was quickly transformed into small sized bubbles, approximately 100 µm size. These bubbles were significantly smaller than those generated in a conventional flotation cell.

The bubble-particle attachment unit was a 0.95-cm-diam stainless steel pipe with a spiral insert which protruded 0.2 cm inward from the pipe wall. The spiral completed one revolution for every 0.5 cm of length throughout the 50-cm-length pipe.

Shallow-depth froth separators of two different sizes were employed. One unit had a 314-cm² surface area and a total volume of 1.7 L. The other unit had a 670-cm² surface area and a total volume of 5.2 L. The bubble-pulp mixture entered the unit through a sparger which was centrally located near the top of the cone shaped tank. The bubbles quickly rose to the surface as the slurry spread out toward the sides of the unit. The froth overflowed at the outer edge and the tailings drained out of the unit through the bottom.

Copper Ore Flotation

The in-line static mixer flotation system was tested using a western porphyry copper ore containing 0.67 pct copper. The goal of the flotation tests was to recover at least 90 pct of the copper from the ore in a concentrate of at least 5 pct copper. A flotation concentrate of 5 pct copper is typical for a rougher product that would be recirculated to the flotation circuit.

The copper ore was obtained from a porphyry copper mine in Arizona. The major sulfide mineral was chalcopyrite with smaller amounts of pyrite. The gangue was mostly plagioclase feldspar with smaller amounts of biotite and magnetite. The chalcopyrite was widely disseminated throughout the ore pieces and the liberation size was approximately 210 µm.

Flotation Procedure

The ore was ground with clear saturated lime water for 10 min in a laboratory rod mill at 70 pct solids and sized at 210 µm. The plus 210-µm size fraction was reground for an additional 8 min. The ground ore pulp was conditioned with 0.05-g/kg potassium amyl xanthate for 10 min in a stirred tank. Dowfroth 10 12 frother was used to stabilize the bubbles. The water to ore ratio in the in-line static mixer was equivalent to 25 pct solids for all the tests. At different aeration levels, the percent solids of the ore slurry was adjusted so that when combined with the bubble slurry, the percent solids in the in-line static mixer was maintained at 25 pct solids. The conditioned sample was placed in the slurry reservoir and diluted to the proper percent solids. The conditioned ore pulp was then pumped through the in-line static mixer along with the bubble slurry to the shallow depth froth separator.

Rapid flotation often requires multiple stages because adequate mineral recovery is not always obtained after one stage. In the second stage, the tailings product from the first stage was dewatered and again fed through the system for the second stage. This sequence was repeated for five stages. Beyond five stages, very little selectivity was obtained which resulted in a small improvement in copper recovery, but a large decrease in the overall concentrate grade. Therefore, for this study, only five flotation stages were used. The concentrates from each stage were combined to form a final concentrate for the multi-stage flotation process. The flotation conditions are summarized in table 1.

Experimental Results and Discussion

Flotation tests were designed to study the effect of mixing intensity in the in-line static mixer, the residence time of the bubbles in the froth separator, and the air to ore ratio. Two different flow rates, 12 and 16 L/min were used to represent different mixing intensities within the in-line static mixer. Two froth separators with surface areas of 314 and 670 cm² were used to study the effect of residence time, Also, air to ore ratios of 0.25, 0.50, and 0.75 mL/g were tested in the in-line static mixer. The operating conditions for the copper flotation tests are shown in table 2.

To determine the flotation rate, the first order kinetics equation was used:

where k is the first order flotation rate constant and tr is the total residence of a particle in the flotation system. The residence time was determined by dividing the total volume of the system by the flow rate. The sum of all the residence times for each flotation stage is included in the total residence time. Since the tailings were dewatered and repumped through the system to simulate flotation staging in a continuous process, the time required for dewatering in the semi-continuous process was not reflected in the overall flotation rate calculation.

Solving the equation for the flotation rate,

The speed of flotation in the different tests would be reflected in the first order flotation rate constant calculated for each test.

Equation 3 was used to calculate the flotation rate for both conventional and rapid flotation tests. Although the two tests were conducted differently, both closely approximated the first order kinetics equation. As shown in figure 3, this provided a means to compare the two flotation systems.

Test results, shown in table 3, show that for both separators and at all three air to ore ratios, the flotation rate and copper recovery were slightly higher at the high flow rate. Using the statistical f-test at a 95 pct confidence level, analysis of variance of the copper recovery showed that the air to ore ratio had a statistically significant effect on the copper recovery, with the highest recoveries obtained at the 0.75 mL/g air to ore ratio. The concentrate grade achieved with the 670-cm² separator was significantly better than the

grade with the smaller separator. The effective bubble residence time in the 670-cm² separator was twice as long as that of the smaller separator. This allowed more of the gangue minerals to drain from the froth which produced a higher grade copper concentrate. The air to ore ratio had the largest effect on the concentrate grade. The best concentrate grades were obtained at 0.5 mL/g air to ore ratio. At 0.75 mL/g air to ore ratio, the flotation rate of the gangue materials also increased which lowered the copper concentrate grade. The flotation goal for this ore was to rapidly float 90 pct of the copper in a high-grade copper concentrate (greater than 5 pct Cu). The highest copper grade was obtained at the 0.50 mL/g air to ore ratio using the 670-cm2 froth separators and the 12-L/min pulp flow rates.

The fastest flotation rate, 2.01 min-¹ (table 4), was obtained at the 12 L/min pulp flow rate with the 314-cm² surface area froth separator. However, the 4.32 pct copper concentrate fell short of the established goal of 5 pct. Therefore, the best recovery (88.3 pct), grade (5.92 pct), and flotation rate (0.89 min-¹) were attained at the 0.50 mL/g air to ore ratio and the 16 L/min pulp flow rate using the 314-cm² froth separator.

A comparison of the flotation results for the rapid flotation system to those of the conventional flotation system is shown in table 5. The rapid flotation system was over seven times faster than the conventional laboratory flotation cell. The capacity (tons per hour per cubic meter) of the rapid flotation system was over seven times greater than that of the conventional system. Therefore, a physically smaller rapid flotation unit could be used to replace a larger conventional flotation unit of the same capacity.

Phosphate Ore Flotation

The in-line static flotation system was tested using a sample of amine phosphate tailings to rapidly float the quartz away from the phosphate. The phosphate-bearing amine phosphate tailings sample, containing 9.5 pct P2O5 in a quartz matrix, was obtained from a Florida phosphate operation. The particle size of the sample ranged between 300 and 38 µm with the P2O5 evenly distributed among the different particle sizes.

Flotation Procedure

The sample was conditioned at pH 6.5 with cornstarch for 30 s to depress the phosphate, and then immediately conditioned with Armac C, an aliphatic amine acetate salt, to float the quartz. Betz M150 frother was used to stabilize the bubbles. The flotation goal was to float a quartz product, low in phosphate, that could be discarded as tailings and to obtain a phosphate – rich product which would be suitable for recycling to the flotation circuit. The water to air ratio in the in-line static mixer was equivalent to 10 pct solids for all the tests. At different aeration levels, the ore slurry percent solids was adjusted so that when combined with the bubble slurry, the percent solids in the in-line static mixer was maintained at 10 percent solids. The conditioned slurry was placed in the slurry reservoir and diluted to the proper percent solids. The conditioned pulp was then pumped through the in-line static mixer along with the bubble slurry. The flotation conditions are summarized in table 6.

Experimental Results and Discussion

Flotation tests were designed to study the effect of mixing intensity in the in-line static mixer, the residence time of the bubbles in the froth separator, and the air to ore ratio. Two different flow rates, 10 and 20 L/min were used to represent different mixing intensities within the in-line static mixer. Two froth separators with surface areas of 314 and 670 cm² were used to study the effect of residence time. Also, air to ore ratios of 1.5, 2.5, and 3.5 mL/g were tested in the in-line static mixer. The air to ore ratio could be adjusted by changing the relative proportions of the ore pulp and bubble slurry being used at each total flow rate. The operating conditions for the quartz flotation tests are shown in table 7.

As in the copper flotation tests, the speed of flotation among different tests would be reflected in the first order flotation rate constant calculated for each test. Although the goal of this flotation process is recovery of phosphate, the flotation rate is a function of the quartz recovery because the quartz is being floated. To determine the flotation rate, again the first order kinetics equation was used.

Test results (table 8), show that for both size separators and at the 1.5 and 2.5 mL/g air to ore ratios, the quartz recovery was higher at the higher flow rate. The air to ore ratio had a statistically significant effect upon the quartz recovery, obtaining the highest recoveries at the 2.5 mL/g air to ore ratio. The best quartz recovery was obtained at the 2.5 mL/g air to ore ratio. The 2.5 mL/g air to ore ratio produced more bubbles than the 1.5 mL/g air to ore ratio, resulting in increased quartz recovery. However, at the 3.5 mL/g air to ore ratio, the quartz recovery declined. This was probably a result of the bubble generator. The bubbles were generated continuously in a constant volume unit and only a portion of the bubble slurry was pumped to the in-line static mixer bubble-particle attachment unit. As the air to ore ratio increased, the flow of water and air to the bubble generator also increased. While the percentage of air in the bubble slurry remained at 35 pct volume, the bubble sizes were visually larger. Less agitation energy was imparted to each unit volume of water and air at this higher rate and the larger bubbles decreased the number of bubble-particle collisions, resulting in a decline in quartz recovery. The flotation rate was affected by both the bubble residence time and the air to ore ratio. At both pulp flow rates and with both size froth separators, the 2.5 mL/g air to ore ratio produced faster flotation rates than the 1.5 mL/g air to ore ratio. The fastest

flotation rate was obtained at the 2.5 mL/g air to ore ratio. Also, as expected, the smaller volume froth separator with its shorter residence time produced the faster flotation rate.

The concentrate grade obtained with the 670-cm² bubble-pulp separator was significantly better than that obtained with the 314-cm² bubble-pulp separator. The effective bubble residence time in the larger separator was twice as long as that in the smaller separator. This allowed more of the phosphate mineral to drain away from the froth; thus, producing a higher grade quartz product. Also, in general the highest grade of the quartz concentrate was obtained at the lower flow rate. At the 10- L/min pulp flow rate, and 2.5 mL/g air to ore ratio, a 90.0-pct silica concentrate was obtained.

While the flotation process recovered a quartz concentrate, the goal of this flotation was to rapidly obtain a high-grade quartz flotation concentrate to discard as a tailings, and to recover most of the phosphate in the unfloated product for recycling to the phosphate flotation circuit. The higher grade quartz products were obtained using either the 1.5 mL/g air to ore ratio at 10-L/min pulp flow rate (48 W mixing intensity) with the 670-cm² surface area froth separator, the 2.5 mL/g air to ore ratio at 10-L/min pulp flow rate (64 W mixing intensity) with the 314-cm² surface area froth separator, or the 2.5 mL/g air to ore ratio at the 10 L/min pulp flow rate (71 W mixing intensity) with the 670-cm² surface area froth separator. The fastest flotation rate among those three conditions was obtained at the 2.5 mL/g air to ore ratio at the 64 W mixing intensity level with the 314-cm² surface area froth separator. About 60 pct of the quartz was rapidly floated at a 5.56-min-¹ flotation rate. The unfloated product contained 17 pct P2O5 and recovered 80 pct of the phosphate. Table 9 compares the results of the best rapid flotation test with the results of the conventional silica flotation test. The grade of the quartz concentrate was the same for both the rapid and conventional flotation systems. The phosphate recovery was higher for the rapid flotation system than for the conventional flotation system, but the phosphate product grade was lower for the rapid flotation system than for the conventional flotation system. The slightly lower grade phosphate product is not a problem because this product will be recirculated to the phosphate recovery circuit. However, the biggest difference between the two techniques was the flotation speed. The flotation rate of the rapid flotation system was over four times faster than that of the conventional flotation system. Since ores processed in the future will be lower grade and grained, larger flotation circuits with much longer residence times will be required for the same production capacity. Therefore, a faster flotation process will be needed to effectively process larger tonnages at smaller particle sizes.

Conclusions

The Bureau-developed rapid flotation system illustrates how the use of discrete units for bubble-particle attachment and bubble-pulp separation can improve the overall flotation kinetics. Each unit can be optimized to obtain the best performance of its function. There is an optimum mixing intensity and slurry residence time in the in-line static mixer for the best performance. There is also an optimum froth separator surface area to flow rate ratio for the froth separator (effective bubble residence time). Matching these two optimized units would substantially improve the flotation kinetics. Improved flotation kinetics allow for much smaller flotation circuits to replace the larger conventional flotation circuits of the same capacity. Using the mixing intensity of the in-line static mixer bubble-particle attachment unit, air to ore ratio, and the froth separator surface area to flow rate ratio parameters, the rapid flotation system could be scaled-up.

The rapid flotation system was shown to be effective on a western porphyry copper ore containing 0.67 pct copper. The best copper grade was obtained at a 0.50 mL/g air to ore ratio, a 16 L/min pulp flow rate, with a 314-cm² surface area froth separator. About 88 pct copper was recovered from the ore in a rougher concentrate containing 5.9 pct copper at a rate over 7 times faster than conventional flotation. The in-line static mixer flotation system was also effective in rapidly floating the quartz away from a phosphate-bearing amine flotation tailings sample. The best quartz grade at the fastest flotation rate was obtained with a 2.5 mL/g air to ore ratio, a 10-L/min pulp flow rate, using a 314-cm² surface area froth separator. Sixty percent of the quartz was rapidly floated at a 5.56-min-¹ flotation rate. This rate was over four times faster than conventional flotation. The unfloated product retained 80 pct of the phosphate in a 16.9-pct concentrate.