cyclone pressure in water balancing - grinding & classification circuits - metallurgist & mineral processing engineer

cyclone pressure in water balancing - grinding & classification circuits - metallurgist & mineral processing engineer

hi guys. i am working at a small scale mining industry. they have been operating at about 80t/h feed to ball mill. the cyclones operates at about 85MPa feed pressure. 2 cyclones run at a time. however they don't maintain water balance .we all know that the principle is "the less the feed the less the water, but proportionally and vice versa". but they keep saying if the water is less proportional to the feed, the cyclone feed pressure will go down, so will result in undesired classification. this is why they add same amount of water all the time whether the feed is less or more. but i suggest, what if we use one cyclone at a time, in case the feed to ball mill goes down from 80t/h to 30t/h. since they use almost same amount of water even though the feed to mill gets low to 30t/h, cyclone overflow becomes diluted and results in a very low density ( it was supposed to be 23%, but goes down to 10-11%), which becomes out of the designed parameter.in my point of view, with the correct water balance i get my cyclone overflow 23% all the time. but how about the cyclone feed pressure? i think we cant increase the pump speed to maintain the pressure, because tonnage to mill is low. The cyclone feed sump can be left empty. here is the help i need guys. Do I have other alternatives to get the designed cyclone pressure or any other ways by keeping cyclone overflow 23% all he time. thanks a lot

I am sure you already know what needs to happen. Any contninous system operates well if it is allowed to attain equilibrium and allowed to continue in equilibrium. any changes that affect equilibrium will affect efficiencies and productivity. one way the deal with is to make sure that you control your feed into the mill from whatever your source is constant, ie.e you need surge capacity before the mill so that feed into you mill is constant as this is critical.

Thanks brighton. Thats a a good idea. But the real problem on the ground is there has been a continuous failure in the crushing section due to liner wear and motor failure, chocking....etc. we fill up the whole stock pile, but the failure on the crusher sometimes lasts 2-3 days. In Which case we will runout of ore if we go with the same tonnage like before at 80-90t/h. This is the problem.

you are totally correct. you want to maintain pressure to achieve same size dist'n. If you add more water and cut cyclone feed density you will make a finer product, dont know if that is good or bad, but downstream density will fall. Large mills have many cyclones in a cyclopac. So removing one is no big deal and gives a stepwise flow control. You are a bit hooped having two cyclones. Your thinking however looks correct

I appreciate that mike. Thanks. We have 6 cyclones. Howevere 2 are running at atime. If I try to add cyclone, my pressure will definitely go down. Its why I choose operating using one cyclone. Our problem is that the size of the apex is static. We cant increase or make it smaller. Nothing automatic means of controlling the apex is attached to it.Of is the same all the time.

The turndown inthroughput is alsolikely having a negative impact on subsequent processing circuits as well.The impact of this instability on their performance probably results inhigher losses (and other measures of performance) during these times.

Thanks Robert that helps a lot. I think I need to see how much money I am going to save or loss, if the mill shut down. i will also have a word with the Grinding Technicians on how bad will that influence on mill operating conditions. Because if this is going to happen, we are going to stop 2-3 days a month. That feels a bit strange right? Thanks again.

Effectively you are right, it is key to work with a balanced water balance to maintain pulp density and grading conditions, defined in the design, but it is also important to maintain the working pressure condition of the cyclone and as it is a equipment that is sizing volumetrically, the correct action is to withdraw equipment from the operation as the feed decreases.Therefore, the action of operating with a cyclone for a feed of 30 t / h is correct, however, the initial feed was 80 t / h, which means two things:1. The mill is being sub-used energetically2. Each cyclone in normal operation classifies a mass flow equivalent to 40 t / h3. A flow of 30 t / h is 25% lower than its normal operating condition4. The above condition will not allow work to the design pressure condition (85 MPa)5. Therefore the classification condition will be deficient and consequently the particle size (P80) for the next stage will be inadequate6. All of the above will negatively impact the metallurgical recovery of the process and the cost of operation (OPEX)7. Increase pumping speed will only exacerbate the problem further, causing problems of accelerated pipe wear and cavitation problems of the pumps when the pumping box is emptied

SMALL CYCLONES:1. Dimension and install a smaller cyclone, suitable for the new feed rate of 30 t / h, if this condition is transient, or replace with a new battery of smaller cyclones, if the condition is permanentBATCH OPERATION:1. It would be advisable to evaluate a batch operation strategy at a processing rate of 80 t / h, starting the operation with the collection of full fine ore (80 x 24 = 1920 tons), which would allow a 64 hour autonomy (2 ,7 days).2. In addition, during the same 64 hours of operation, an additional 1920 tons of ore could be collected (30 x 64 = 1920 tons), which would achieve a further 64 hours of autonomy3. The proposed strategy would allow a total autonomy of 128 hours (5.3 days) operating at 80 t / h with 64 hours (2.7 days) of detention, intended for maintenance and replenishment of fine ore collection, among other activities4. To clear any doubts about the advantages and disadvantages of both options, it would be advisable to conduct a trade off, comparing energy consumption, energy efficiency, availability, utilization and opex.

To maintain the pressure at the hydrocyclone inlet , you should maintain the % solids in the range of 45-55 % at the hydrocyclone inlet and it can be very easily monitored ene . Densitometer will sense the pulp density and based on that it will operate the water control valve at the ball mill outlet and in that way you can automatically control the Feed parameters . This is a permanent solution to your problem .

Thanks for the suggestion ashutosh. But I dont think u understand the problem. My cyclone feed percent is 55%. But that works out when feed is 90t/h. If feed goes down to 40 t/h , you cant do the correct water balance and continue with 55%. The pump is a cyclone pump. Its impeller is big. You try to maintain the pressure through the velocity of the impeller as usual, your sump will be left empty sooner. the pump will suck air and cause problem. I dont want this to happen.this is why I ask the professionals for other options of there is any. I in the comment I wrote before, my apex is static. If it can be altered, it was the best shot. Anyways Thanks for the suggestion.

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henan mining machinery and equipment manufacturer - what does a cyclone do in the gold mining process

henan mining machinery and equipment manufacturer - what does a cyclone do in the gold mining process

They are literally a gold mine of placer ... and obtain good results in gold recovery. Dry processing recovery systems generally use air flows to do the same job that ...Gold Mining Process Development. THE BASIC PROCESSES OF GOLD RECOVERY INTRODUCTION. ... The flotation process in general does not float free gold particles but is

Ore beneficiation equipment, sand making equipment, crushing equipment and powder grinding equipment, which are widely used in various industries such as metallurgy, mine, chemistry, building material, coal, refractory and ceramics.

ball mill - an overview | sciencedirect topics

ball mill - an overview | sciencedirect topics

The ball mill accepts the SAG or AG mill product. Ball mills give a controlled final grind and produce flotation feed of a uniform size. Ball mills tumble iron or steel balls with the ore. The balls are initially 510 cm diameter but gradually wear away as grinding of the ore proceeds. The feed to ball mills (dry basis) is typically 75 vol.-% ore and 25% steel.

The ball mill is operated in closed circuit with a particle-size measurement device and size-control cyclones. The cyclones send correct-size material on to flotation and direct oversize material back to the ball mill for further grinding.

Grinding elements in ball mills travel at different velocities. Therefore, collision force, direction and kinetic energy between two or more elements vary greatly within the ball charge. Frictional wear or rubbing forces act on the particles, as well as collision energy. These forces are derived from the rotational motion of the balls and movement of particles within the mill and contact zones of colliding balls.

By rotation of the mill body, due to friction between mill wall and balls, the latter rise in the direction of rotation till a helix angle does not exceed the angle of repose, whereupon, the balls roll down. Increasing of rotation rate leads to growth of the centrifugal force and the helix angle increases, correspondingly, till the component of weight strength of balls become larger than the centrifugal force. From this moment the balls are beginning to fall down, describing during falling certain parabolic curves (Figure 2.7). With the further increase of rotation rate, the centrifugal force may become so large that balls will turn together with the mill body without falling down. The critical speed n (rpm) when the balls are attached to the wall due to centrifugation:

where Dm is the mill diameter in meters. The optimum rotational speed is usually set at 6580% of the critical speed. These data are approximate and may not be valid for metal particles that tend to agglomerate by welding.

The degree of filling the mill with balls also influences productivity of the mill and milling efficiency. With excessive filling, the rising balls collide with falling ones. Generally, filling the mill by balls must not exceed 3035% of its volume.

The mill productivity also depends on many other factors: physical-chemical properties of feed material, filling of the mill by balls and their sizes, armor surface shape, speed of rotation, milling fineness and timely moving off of ground product.

where b.ap is the apparent density of the balls; l is the degree of filling of the mill by balls; n is revolutions per minute; 1, and 2 are coefficients of efficiency of electric engine and drive, respectively.

A feature of ball mills is their high specific energy consumption; a mill filled with balls, working idle, consumes approximately as much energy as at full-scale capacity, i.e. during grinding of material. Therefore, it is most disadvantageous to use a ball mill at less than full capacity.

The ball mill is a tumbling mill that uses steel balls as the grinding media. The length of the cylindrical shell is usually 11.5 times the shell diameter (Figure 8.11). The feed can be dry, with less than 3% moisture to minimize ball coating, or slurry containing 2040% water by weight. Ball mills are employed in either primary or secondary grinding applications. In primary applications, they receive their feed from crushers, and in secondary applications, they receive their feed from rod mills, AG mills, or SAG mills.

Ball mills are filled up to 40% with steel balls (with 3080mm diameter), which effectively grind the ore. The material that is to be ground fills the voids between the balls. The tumbling balls capture the particles in ball/ball or ball/liner events and load them to the point of fracture.

When hard pebbles rather than steel balls are used for the grinding media, the mills are known as pebble mills. As mentioned earlier, pebble mills are widely used in the North American taconite iron ore operations. Since the weight of pebbles per unit volume is 3555% of that of steel balls, and as the power input is directly proportional to the volume weight of the grinding medium, the power input and capacity of pebble mills are correspondingly lower. Thus, in a given grinding circuit, for a certain feed rate, a pebble mill would be much larger than a ball mill, with correspondingly a higher capital cost. However, the increase in capital cost is justified economically by a reduction in operating cost attributed to the elimination of steel grinding media.

In general, ball mills can be operated either wet or dry and are capable of producing products in the order of 100m. This represents reduction ratios of as great as 100. Very large tonnages can be ground with these ball mills because they are very effective material handling devices. Ball mills are rated by power rather than capacity. Today, the largest ball mill in operation is 8.53m diameter and 13.41m long with a corresponding motor power of 22MW (Toromocho, private communications).

Planetary ball mills. A planetary ball mill consists of at least one grinding jar, which is arranged eccentrically on a so-called sun wheel. The direction of movement of the sun wheel is opposite to that of the grinding jars according to a fixed ratio. The grinding balls in the grinding jars are subjected to superimposed rotational movements. The jars are moved around their own axis and, in the opposite direction, around the axis of the sun wheel at uniform speed and uniform rotation ratios. The result is that the superimposition of the centrifugal forces changes constantly (Coriolis motion). The grinding balls describe a semicircular movement, separate from the inside wall, and collide with the opposite surface at high impact energy. The difference in speeds produces an interaction between frictional and impact forces, which releases high dynamic energies. The interplay between these forces produces the high and very effective degree of size reduction of the planetary ball mill. Planetary ball mills are smaller than common ball mills, and are mainly used in laboratories for grinding sample material down to very small sizes.

Vibration mill. Twin- and three-tube vibrating mills are driven by an unbalanced drive. The entire filling of the grinding cylinders, which comprises the grinding media and the feed material, constantly receives impulses from the circular vibrations in the body of the mill. The grinding action itself is produced by the rotation of the grinding media in the opposite direction to the driving rotation and by continuous head-on collisions of the grinding media. The residence time of the material contained in the grinding cylinders is determined by the quantity of the flowing material. The residence time can also be influenced by using damming devices. The sample passes through the grinding cylinders in a helical curve and slides down from the inflow to the outflow. The high degree of fineness achieved is the result of this long grinding procedure. Continuous feeding is carried out by vibrating feeders, rotary valves, or conveyor screws. The product is subsequently conveyed either pneumatically or mechanically. They are basically used to homogenize food and feed.

CryoGrinder. As small samples (100 mg or <20 ml) are difficult to recover from a standard mortar and pestle, the CryoGrinder serves as an alternative. The CryoGrinder is a miniature mortar shaped as a small well and a tightly fitting pestle. The CryoGrinder is prechilled, then samples are added to the well and ground by a handheld cordless screwdriver. The homogenization and collection of the sample is highly efficient. In environmental analysis, this system is used when very small samples are available, such as small organisms or organs (brains, hepatopancreas, etc.).

The vibratory ball mill is another kind of high-energy ball mill that is used mainly for preparing amorphous alloys. The vials capacities in the vibratory mills are smaller (about 10 ml in volume) compared to the previous types of mills. In this mill, the charge of the powder and milling tools are agitated in three perpendicular directions (Fig. 1.6) at very high speed, as high as 1200 rpm.

Another type of the vibratory ball mill, which is used at the van der Waals-Zeeman Laboratory, consists of a stainless steel vial with a hardened steel bottom, and a single hardened steel ball of 6 cm in diameter (Fig. 1.7).

The mill is evacuated during milling to a pressure of 106 Torr, in order to avoid reactions with a gas atmosphere.[44] Subsequently, this mill is suitable for mechanical alloying of some special systems that are highly reactive with the surrounding atmosphere, such as rare earth elements.

A ball mill is a relatively simple apparatus in which the motion of the reactor, or of a part of it, induces a series of collisions of balls with each other and with the reactor walls (Suryanarayana, 2001). At each collision, a fraction of the powder inside the reactor is trapped between the colliding surfaces of the milling tools and submitted to a mechanical load at relatively high strain rates (Suryanarayana, 2001). This load generates a local nonhydrostatic mechanical stress at every point of contact between any pair of powder particles. The specific features of the deformation processes induced by these stresses depend on the intensity of the mechanical stresses themselves, on the details of the powder particle arrangement, that is on the topology of the contact network, and on the physical and chemical properties of powders (Martin et al., 2003; Delogu, 2008a). At the end of any given collision event, the powder that has been trapped is remixed with the powder that has not undergone this process. Correspondingly, at any instant in the mechanical processing, the whole powder charge includes fractions of powder that have undergone a different number of collisions.

The individual reactive processes at the perturbed interface between metallic elements are expected to occur on timescales that are, at most, comparable with the collision duration (Hammerberg et al., 1998; Urakaev and Boldyrev, 2000; Lund and Schuh, 2003; Delogu and Cocco, 2005a,b). Therefore, unless the ball mill is characterized by unusually high rates of powder mixing and frequency of collisions, reactive events initiated by local deformation processes at a given collision are not affected by a successive collision. Indeed, the time interval between successive collisions is significantly longer than the time period required by local structural perturbations for full relaxation (Hammerberg et al., 1998; Urakaev and Boldyrev, 2000; Lund and Schuh, 2003; Delogu and Cocco, 2005a,b).

These few considerations suffice to point out the two fundamental features of powder processing by ball milling, which in turn govern the MA processes in ball mills. First, mechanical processing by ball milling is a discrete processing method. Second, it has statistical character. All of this has important consequences for the study of the kinetics of MA processes. The fact that local deformation events are connected to individual collisions suggests that absolute time is not an appropriate reference quantity to describe mechanically induced phase transformations. Such a description should rather be made as a function of the number of collisions (Delogu et al., 2004). A satisfactory description of the MA kinetics must also account for the intrinsic statistical character of powder processing by ball milling. The amount of powder trapped in any given collision, at the end of collision is indeed substantially remixed with the other powder in the reactor. It follows that the same amount, or a fraction of it, could at least in principle be trapped again in the successive collision.

This is undoubtedly a difficult aspect to take into account in a mathematical description of MA kinetics. There are at least two extreme cases to consider. On the one hand, it could be assumed that the powder trapped in a given collision cannot be trapped in the successive one. On the other, it could be assumed that powder mixing is ideal and that the amount of powder trapped at a given collision has the same probability of being processed in the successive collision. Both these cases allow the development of a mathematical model able to describe the relationship between apparent kinetics and individual collision events. However, the latter assumption seems to be more reliable than the former one, at least for commercial mills characterized by relatively complex displacement in the reactor (Manai et al., 2001, 2004).

A further obvious condition for the successful development of a mathematical description of MA processes is the one related to the uniformity of collision regimes. More specifically, it is highly desirable that the powders trapped at impact always experience the same conditions. This requires the control of the ball dynamics inside the reactor, which can be approximately obtained by using a single milling ball and an amount of powder large enough to assure inelastic impact conditions (Manai et al., 2001, 2004; Delogu et al., 2004). In fact, the use of a single milling ball avoids impacts between balls, which have a remarkable disordering effect on the ball dynamics, whereas inelastic impact conditions permit the establishment of regular and periodic ball dynamics (Manai et al., 2001, 2004; Delogu et al., 2004).

All of the above assumptions and observations represent the basis and guidelines for the development of the mathematical model briefly outlined in the following. It has been successfully applied to the case of a Spex Mixer/ Mill mod. 8000, but the same approach can, in principle, be used for other ball mills.

The Planetary ball mills are the most popular mills used in MM, MA, and MD scientific researches for synthesizing almost all of the materials presented in Figure 1.1. In this type of mill, the milling media have considerably high energy, because milling stock and balls come off the inner wall of the vial (milling bowl or vial) and the effective centrifugal force reaches up to 20 times gravitational acceleration.

The centrifugal forces caused by the rotation of the supporting disc and autonomous turning of the vial act on the milling charge (balls and powders). Since the turning directions of the supporting disc and the vial are opposite, the centrifugal forces alternately are synchronized and opposite. Therefore, the milling media and the charged powders alternatively roll on the inner wall of the vial, and are lifted and thrown off across the bowl at high speed, as schematically presented in Figure 2.17.

However, there are some companies in the world who manufacture and sell number of planetary-type ball mills; Fritsch GmbH (www.fritsch-milling.com) and Retsch (http://www.retsch.com) are considered to be the oldest and principal companies in this area.

Fritsch produces different types of planetary ball mills with different capacities and rotation speeds. Perhaps, Fritsch Pulverisette P5 (Figure 2.18(a)) and Fritsch Pulverisette P6 (Figure 2.18(b)) are the most popular models of Fritsch planetary ball mills. A variety of vials and balls made of different materials with different capacities, starting from 80ml up to 500ml, are available for the Fritsch Pulverisette planetary ball mills; these include tempered steel, stainless steel, tungsten carbide, agate, sintered corundum, silicon nitride, and zirconium oxide. Figure 2.19 presents 80ml-tempered steel vial (a) and 500ml-agate vials (b) together with their milling media that are made of the same materials.

Figure 2.18. Photographs of Fritsch planetary-type high-energy ball mill of (a) Pulverisette P5 and (b) Pulverisette P6. The equipment is housed in the Nanotechnology Laboratory, Energy and Building Research Center (EBRC), Kuwait Institute for Scientific Research (KISR).

Figure 2.19. Photographs of the vials used for Fritsch planetary ball mills with capacity of (a) 80ml and (b) 500ml. The vials and the balls shown in (a) and (b) are made of tempered steel agate materials, respectively (Nanotechnology Laboratory, Energy and Building Research Center (EBRC), Kuwait Institute for Scientific Research (KISR)).

More recently and in year 2011, Fritsch GmbH (http://www.fritsch-milling.com) introduced a new high-speed and versatile planetary ball mill called Planetary Micro Mill PULVERISETTE 7 (Figure 2.20). The company claims this new ball mill will be helpful to enable extreme high-energy ball milling at rotational speed reaching to 1,100rpm. This allows the new mill to achieve sensational centrifugal accelerations up to 95 times Earth gravity. They also mentioned that the energy application resulted from this new machine is about 150% greater than the classic planetary mills. Accordingly, it is expected that this new milling machine will enable the researchers to get their milled powders in short ball-milling time with fine powder particle sizes that can reach to be less than 1m in diameter. The vials available for this new type of mill have sizes of 20, 45, and 80ml. Both the vials and balls can be made of the same materials, which are used in the manufacture of large vials used for the classic Fritsch planetary ball mills, as shown in the previous text.

Retsch has also produced a number of capable high-energy planetary ball mills with different capacities (http://www.retsch.com/products/milling/planetary-ball-mills/); namely Planetary Ball Mill PM 100 (Figure 2.21(a)), Planetary Ball Mill PM 100 CM, Planetary Ball Mill PM 200, and Planetary Ball Mill PM 400 (Figure 2.21(b)). Like Fritsch, Retsch offers high-quality ball-milling vials with different capacities (12, 25, 50, 50, 125, 250, and 500ml) and balls of different diameters (540mm), as exemplified in Figure 2.22. These milling tools can be made of hardened steel as well as other different materials such as carbides, nitrides, and oxides.

Figure 2.21. Photographs of Retsch planetary-type high-energy ball mill of (a) PM 100 and (b) PM 400. The equipment is housed in the Nanotechnology Laboratory, Energy and Building Research Center (EBRC), Kuwait Institute for Scientific Research (KISR).

Figure 2.22. Photographs of the vials used for Retsch planetary ball mills with capacity of (a) 80ml, (b) 250ml, and (c) 500ml. The vials and the balls shown are made of tempered steel (Nanotechnology Laboratory, Energy and Building Research Center (EBRC), Kuwait Institute for Scientific Research (KISR)).

Both Fritsch and Retsch companies have offered special types of vials that allow monitoring and measure the gas pressure and temperature inside the vial during the high-energy planetary ball-milling process. Moreover, these vials allow milling the powders under inert (e.g., argon or helium) or reactive gas (e.g., hydrogen or nitrogen) with a maximum gas pressure of 500kPa (5bar). It is worth mentioning here that such a development made on the vials design allows the users and researchers to monitor the progress tackled during the MA and MD processes by following up the phase transformations and heat realizing upon RBM, where the interaction of the gas used with the freshly created surfaces of the powders during milling (adsorption, absorption, desorption, and decomposition) can be monitored. Furthermore, the data of the temperature and pressure driven upon using this system is very helpful when the ball mills are used for the formation of stable (e.g., intermetallic compounds) and metastable (e.g., amorphous and nanocrystalline materials) phases. In addition, measuring the vial temperature during blank (without samples) high-energy ball mill can be used as an indication to realize the effects of friction, impact, and conversion processes.

More recently, Evico-magnetics (www.evico-magnetics.de) has manufactured an extraordinary high-pressure milling vial with gas-temperature-monitoring (GTM) system. Likewise both system produced by Fritsch and Retsch, the developed system produced by Evico-magnetics, allowing RBM but at very high gas pressure that can reach to 15,000kPa (150bar). In addition, it allows in situ monitoring of temperature and of pressure by incorporating GTM. The vials, which can be used with any planetary mills, are made of hardened steel with capacity up to 220ml. The manufacturer offers also two-channel system for simultaneous use of two milling vials.

Using different ball mills as examples, it has been shown that, on the basis of the theory of glancing collision of rigid bodies, the theoretical calculation of tPT conditions and the kinetics of mechanochemical processes are possible for the reactors that are intended to perform different physicochemical processes during mechanical treatment of solids. According to the calculations, the physicochemical effect of mechanochemical reactors is due to short-time impulses of pressure (P = ~ 10101011 dyn cm2) with shift, and temperature T(x, t). The highest temperature impulse T ~ 103 K are caused by the dry friction phenomenon.

Typical spatial and time parameters of the impactfriction interaction of the particles with a size R ~ 104 cm are as follows: localization region, x ~ 106 cm; time, t ~ 108 s. On the basis of the obtained theoretical results, the effect of short-time contact fusion of particles treated in various comminuting devices can play a key role in the mechanism of activation and chemical reactions for wide range of mechanochemical processes. This role involves several aspects, that is, the very fact of contact fusion transforms the solid phase process onto another qualitative level, judging from the mass transfer coefficients. The spatial and time characteristics of the fused zone are such that quenching of non-equilibrium defects and intermediate products of chemical reactions occurs; solidification of the fused zone near the contact point results in the formation of a nanocrystal or nanoamor- phous state. The calculation models considered above and the kinetic equations obtained using them allow quantitative ab initio estimates of rate constants to be performed for any specific processes of mechanical activation and chemical transformation of the substances in ball mills.

There are two classes of ball mills: planetary and mixer (also called swing) mill. The terms high-speed vibration milling (HSVM), high-speed ball milling (HSBM), and planetary ball mill (PBM) are often used. The commercial apparatus are PBMs Fritsch P-5 and Fritsch Pulverisettes 6 and 7 classic line, the Retsch shaker (or mixer) mills ZM1, MM200, MM400, AS200, the Spex 8000, 6750 freezer/mill SPEX CertiPrep, and the SWH-0.4 vibrational ball mill. In some instances temperature controlled apparatus were used (58MI1); freezer/mills were used in some rare cases (13MOP1824).

The balls are made of stainless steel, agate (SiO2), zirconium oxide (ZrO2), or silicon nitride (Si3N). The use of stainless steel will contaminate the samples with steel particles and this is a problem both for solid-state NMR and for drug purity.

However, there are many types of ball mills (see Chapter 2 for more details), such as drum ball mills, jet ball mills, bead-mills, roller ball mills, vibration ball mills, and planetary ball mills, they can be grouped or classified into two types according to their rotation speed, as follows: (i) high-energy ball mills and (ii) low-energy ball mills. Table 3.1 presents characteristics and comparison between three types of ball mills (attritors, vibratory mills, planetary ball mills and roller mills) that are intensively used on MA, MD, and MM techniques.

In fact, choosing the right ball mill depends on the objectives of the process and the sort of materials (hard, brittle, ductile, etc.) that will be subjecting to the ball-milling process. For example, the characteristics and properties of those ball mills used for reduction in the particle size of the starting materials via top-down approach, or so-called mechanical milling (MM process), or for mechanically induced solid-state mixing for fabrications of composite and nanocomposite powders may differ widely from those mills used for achieving mechanically induced solid-state reaction (MISSR) between the starting reactant materials of elemental powders (MA process), or for tackling dramatic phase transformation changes on the structure of the starting materials (MD). Most of the ball mills in the market can be employed for different purposes and for preparing of wide range of new materials.

Martinez-Sanchez et al. [4] have pointed out that employing of high-energy ball mills not only contaminates the milled amorphous powders with significant volume fractions of impurities that come from milling media that move at high velocity, but it also affects the stability and crystallization properties of the formed amorphous phase. They have proved that the properties of the formed amorphous phase (Mo53Ni47) powder depends on the type of the ball-mill equipment (SPEX 8000D Mixer/Mill and Zoz Simoloter mill) used in their important investigations. This was indicated by the high contamination content of oxygen on the amorphous powders prepared by SPEX 8000D Mixer/Mill, when compared with the corresponding amorphous powders prepared by Zoz Simoloter mill. Accordingly, they have attributed the poor stabilities, indexed by the crystallization temperature of the amorphous phase formed by SPEX 8000D Mixer/Mill to the presence of foreign matter (impurities).

grinding mill - an overview | sciencedirect topics

grinding mill - an overview | sciencedirect topics

The principle objective for controlling grinding mill operation is to produce a product having an acceptable and constant size distribution at optimum cost. To achieve this objective an attempt is made to stabilize the operation by principally controlling the process variables. The main disturbances in a grinding circuit are:

The mill control strategy has to compensate for these variations and minimize any disturbances to the hydrocyclone that is usually in closed circuit. The simplest arrangement is to setup several control loops starting from the control of water/solid ratio in the feed slurry, sump level control, density control of pulp streams at various stages and control of circulating load. Presently most mills use centrifugal pumps for discharging from the sump. This helps to counter surges and other problems related to pumping. For feed control the most likely option is to use a feed forward control while for controlling the hopper level and mill speed and other loops the PI or PID controller is used. The control action should be fast enough to prevent the sump from overflowing or drying out. This can be attained by a cascade control system. The set point of the controller is determined from the level control loop. This type of control promotes stability.

Each of these is controlled by specific controlled inputs, i.e., feed rate, feed water and discharge water flows. The overflow solids fraction is controlled by monitoring the ratio of total water addition (WTOT) to the solid feed rate. The ratio being fixed by the target set point of the overflow solid fraction.

Usually the charge volume of SAG mills occupy between 30-40% of its internal volume at which the grinding rate is maximized. When the charge volume is more, then the throughput suffers. The fill level is monitored by mill weight measurement as most modem mills are invariably mounted on load-cells.

During the operation of SAG mills, it is sometimes observed that the sump levels fall sharply and so does the power draft. This phenomenon is attributed to flow restrictions against the grate. When this occurs it is necessary to control, (or in extreme circumstances), stop the incoming feed.

The power draft is the result of the torque produced by the mill charge density, lift angle of the charge within the mill and fill level. The relationship between these parameters is complex and difficult. Therefore to control mill operation by power draft alone is difficult.

For the purpose of stabilization of the circuit, the basis is to counteract the disturbances. Also the set points must be held. The set points are attributed by dynamic mass balances at each stage of the circuit.

In modern practice the structure and instrumentation of the control systems of tubular grinding mills are designed to operate in three levels or in some cases four levels. The control loops and sensors for a SAG-mill and the levels of control are illustrated in Fig. 18.22. According to Elber [10,17], the levels are:

The function of Level 2 is to stabilize the circuit and to provide the basis of optimizing function in Level 3. Three cascade loops operating in level 2 controls that function in conjunction with level 1 controllers. The cascade loops are:

The set points are supplied by level 3 controllers for all the cascade loops. The mill load and percent solids in the two streams are calculated from signals received by sensors in the water flow stream, the sump discharge flow rate and the density readings from density meters in the pulp streams. The mill load cells supply the charge mass. The load cell signals are compensated for pinion up thrusts [10].

To determine the set point for the optimum mill load, a relation between load, consisting of different feed blends and performance (the maximum achievable throughput) is established. Similar observations are made for mill discharge density and mill discharge flow.

The primary function at Level 3 is optimisation of the SAG mill operation. That is, control of the product at optimum level. In an integrated situation where ball mill and cyclone is in the circuit, the optimisation must take place keeping in mind the restraints imposed by down stream requirements. This optimisation can best be achieved by developing a software for computer use. Usually a large database is required to cover infrequent control actions.

Ore samples were taken from a grinding mill operating as a batch process. The feed size distribution, breakage functions and size analysis of samples taken at intervals of 10minutes up to 30minutes are given. Determine:

An alkaline slurry from a bauxite grinding mill was scheduled to be classified using a spiral classifier at the underflow rate of 1100t/day. The width of the classifier flight was 1.3m and the outside diameter of the spiral flights was 1.2m. Estimate the pitch of the spirals if the spiral speed is 5rev/min and the bulk density of the underflow solids is 2000kg/m3.

The diameter of a typical hydrocyclone was 30.5cm. The apex was fitted with a rubber sleeve 12cm in length and 8.0cm in internal diameter. A quartz suspension at a density of 1.33 was fed to the cyclone at the rate of 1000L/min. The underflow measured 75% solids. The apex diameter was reduced by 10% twice. Estimate:

The volume flow rate of pulp fed to a hydrocyclone was 129L/min. Its solid content was held at 32% by volume. Samples of the feed, under flow and over flow streams were taken simultaneously, dried and a size analysis carried out. The results obtained were:

An hydrocyclone is to be installed in a closed circuit grinding circuit with a mill discharge containing 30% solids by volume. The solid density is 2800kg/m3 and the density of water is 1000kg/m3. Given that the maximum pressure deferential between the inlet and overflow was 50kPa and the throughput from the mill was 800t/h, estimate:

A hydrocyclone classifier is fed with quartz slurry at the rate of 20.8t/h from a grinding mill. The underflow is recirculated. The screen analysis of each stream were determined with the following results:

The input and output streams of an operating cyclone were sampled simultaneously for the same period of time. The dried samples were analysed for size distribution and the mass per cent retained on each size fraction was determined with the following results:

After a steady state operation the solid content of feed slurry was increased by 20% while all other conditions remained the same. Determine the size distribution of each stream under the altered condition.

If a second cyclone is added in series to the cyclone in problem 12.8, what is the effect of the overall efficiency of the classification. What will be the size distribution of the cyclone U/F of the second stage? The partition coefficient of the second stage cyclone is given as:

A crushing plant delivered ore to a wet grinding mill for further size reduction. The size of crushed ore (F80) was. 4.0mm and the S.G. 2.8t/m3. The work index of the ore was determined as 12.2kWh/t. A wet ball mill 1m1m was chosen to grind the ore down to 200microns. A 30% pulp was made and charged to the mill, which was then rotated at 60% of the critical speed. Estimate:

Everell [34] believed that the mechanism of breakage of particles in a grinding mill was analogous to the slow compression loading of irregular particles and that the specific rate of breakage for a particular size of fragment is an inverse function of the average failure load of the particles. Everell et al [35] developed a model to describe the relationship between the grinding selection function (breakage rate) and the physio-mechanical properties of the rocks. The advantage in such a relationship lies in the wealth of rock strength data determined on drill core during mine development being available to predict energy demands in the comminution circuits.

Briggs [36] measured the tensile strength, using the Brazil tensile test, and the point load compressive strength of four rock types of different grindabilities. These results were compared to the Bond Work Index of the ores as measured by the Magdalinovic method [27]. The results in Fig. 3.8 show that there is a good correlation between the Bond Work Index and the tensile strength and the Equivalent Uniaxial Compressive Strength (EUCS). Some of the scatter in the graphs are due to the structure of the rock. For example one rock type was a banded iron, heavily mineralised with sulphides with numerous planes of weakness on a macro scale. This affected the mechanical properties when tested on large specimens. However when the grinding tests were carried out at relatively small particle sizes, the planes of weakness were no longer present and the ore became more competent.

The correlation between Bond Work Index and tensile strength is an indication that the grinding mechanism in the work index test favours abrasion type breakage given that tensile strength is a fair indication of the abrasiveness of a rock.

Briggs also measured the breakage rate and breakage distribution function of the ores and compared the breakage rate and Bond Work Index. There was a good correlation between the rock strength data and the breakage rate with higher strength rocks having a lower breakage rate. However the data set for these tests was small and further work needs to be done to confirm the relationship.

Bearman et al [38] measured a wide range of rock strength properties and correlated these to the JKMRC drop weight test data. Conclusions were that this technique will enable the data required for comminution plant design to be obtained from mechanical tests on drill core samples.

During normal operation the mill speed tends to vary with mill charge. According to available literature, the operating speeds of AG mills are much higher than conventional tumbling mills and are in the range of 8085% of the critical speed. SAG mills of comparable size but containing say 10% ball charge (in addition to the rocks), normally operate between 70 to 75% of the critical speed. Dry Aerofall mills are run at about 85% of the critical speed.

The breakage of particles depends on the speed of rotation. Working with a 7.32m diameter and 3.66m long mill Napier-Munn et al [4] observed that the breakage rate for the finer size fractions of ore (say 0.1mm) at lower speeds (eg. 55% of the critical speed), were higher than that observed at higher speeds (eg. 70% of the critical speed). For larger sizes of ore, (in excess of 10mm), the breakage rate was lower for mills rotating at 55% of the critical speed than for mills running at 70% of the critical speed. For a particular intermediate particle size range, indications are that the breakage rate was independent of speed. The breakage ratesize relation at two different speeds is reproduced in Fig. 9.7.

Ultrafine grinding (UFG) has continued to evolve in terms of equipment development. A number of specialist machines are commercially available including Xstrata's IsaMill, Metso's Vertimill, Outotec's High Intensity Grinding (HIG) mill, and the Metprotech mill. UFG equipment has been developed with installed powers of up to 5MW.

Compared with conventional ball or pebble milling, the specialist machines are significantly more energy efficient and can economically grind to 10m or lower, whereas the economical limit on conventional regrind mills was generally considered to be around 30m. Coupled with improvements in downstream flotation and oxidation processes, the rise of UFG has enabled treatment of more finely grained refractory ores due to a higher degree of liberation in the case of flotation or enhanced oxidation due to the generation of higher surface areas.

In 1993, the Salsigne Gold Mine was reopened. Salsigne treated a gold-bearing pyrite/arsenopyrite ore by flotation, with the flotation tails treated in a CIL circuit and the concentrate reground in a conventional mill to approximately 2530m. The oxygen demand for reground concentrate was high and the rate of oxidation was slow. The concentrate was initially oxidized for approximately 6h using oxygen injection via a Filblast aerator before cyanidation. Additional oxygen was added in in the second CIL stage and hydrogen peroxide was added into the fourth unit to maintain dissolved oxygen concentrations of >10ppm.

Goldcorp have commenced operations at the Elenore Gold Project in Quebec, Canada. The mineralogy of the ore and hence the circuit selection show similarities to those at Salsigne. The main sulfides are arsenopyrite, pyrite, and pyrhottite. The ore is floated, with the flotation tails passing to a tails CIL circuit and the flotation concentrate reground before passing to the concentrate CIL circuit via preaeration tanks designed to achieve 18-h contact with oxygen. The main difference between the Salsigne and Elenore projects is that the Elenore concentrate is ground to 10m and oxidation of the sulfides is substantially complete before cyanidation.

Large projects are typically associated with more complex contracting strategies but not necessarily greater flow sheet complexity. The complexity of the contracting strategy and the increased focus on key items of equipment, such as large grinding mills, elevate the manning requirements. Higher throughput, and associated larger equipment, does lead to increased complexity in service equipment such as lubrication, cooling, and control systems. Gearless motor drives on large semi-autogenous grinding (SAG) and ball mills require significant installation testing and commissioning effort. The 20 MW drive for a large Australian gold/copper project with a capital cost of approximately A$295M in 1998, took three technicians over 6weeks to test and commission with the total vendor cost (installation and commissioning) for the 20MW ring motor alone exceeding A$1M.

The first step in slag processing is size reduction to liberate metallic iron and iron-bearing minerals. This is done by crushers or by autogenous grinding, that is, the slag is ground on its own in the grinding mill without any balls. The latter process yields higher quality product as the iron product discharged from grinding mill contains as high as 80% Fe (Shen and Forssberg, 2003). Metallic iron and iron minerals are separated by magnetic separation. The phosphorus-bearing minerals occurring in steel slag are removed in the tailings of high gradient magnetic separation. The flow diagram is shown in Figure 8.2.

Reduction of iron oxide at high temperature has been shown to be an attractive low energy cost process (Olginskij and Prokhorenko, 1994). The iron-free mineral residue is suitable for applications in construction industry. An alternative route applies microwave heating with carbon and the recovery of iron by magnetic separation (Hatton and Pickles, 1994).

Low-competency ores such as oxides are unlikely to have a problem with generation of pebbles. They are more likely to have a problem with slurry viscosity. For smaller plant in the 1980s, these ores were treated through a single-stage SAG mill grinding to 75 or 106m. This type of circuit is still straightforward in its design concept. High-competency ores or ores requiring a finer product size frequently require two stages of grinding and a number of design issues become important.

Mill orientation. For smaller plant, the mills can be arranged in parallel if the products feed the same discharge hopper. This is more difficult for larger mills as the diameter of the mills drives the height required to maintain the discharge launder slope. In this situation the mills are often located at right angles. Alternatively, a transfer hopper can be used. This decision is usually based on the difference in cost and operability of each option.

Transfer size from the SAG to the ball mill. Transfer size prediction is somewhat uncertain but has an important impact on the balance of power between the two stages of the milling circuit. For feasibility studies, a combination of modeling and benchmarking is generally used.

Mill discharge arrangement. If the SAG and ball mills feed separate pump systems this is not an issue. If the two stages feed the same pump system, the SAG mill usually drives the discharge arrangement. This is because the SAG mill is usually of larger diameter and may have a requirement for screening of pebbles to protect the mill discharge pumps and cyclones, and to facilitate pebble crushing.

how coarse can ball mill feed be - grinding & classification circuits - metallurgist & mineral processing engineer

how coarse can ball mill feed be - grinding & classification circuits - metallurgist & mineral processing engineer

We are looking at a 1 MTPA comminution circuit that goes from a tertiary crush to a single ball mill (closed on a cyclone cluster) to produce a 100 m P80 product. We're doing a trade-off study. Closing the tertiary on the screen or leaving it open! (The crushing plant runs 11 h per day, the mill 24 h/d.) The mill power (and media consumption) is higher and the mill grinding length is a bit longer in the open case. Is there any other side-effects associated with feeding a ball mill with 16 kW-h/tonne feed which is coarser (closed P80 of 13.4 mm vs. open P80 of 15.4 mm i.e.2 mm coarser)? Aside from the additional capital and operating costs are there any operational issues associated with feed which is coarse i.e. coarser than the selection function maximum which is at -2 mm.

High scatting rate would be the primary concern I think, and as the crusher liners wear, the situation will get worse and your throughput rate/grind size will vary. Wouldn't recommend it if this is going to be an issue.

What we're looking at is a jaw, open circuit secondary cone and an open circuit tertiary. Screens before the cones if the capacity is a problem but no recycle on the secondary or tertiary. The idea is the very cheapest get-into-operation solution. The cost of the recycle conveyors and larger screens is being avoided (and being transferred to mill op costs). What I really am interested in is whether the scatting problem is a definite or a maybe if the Bond Ball Mill Work Index were 16 kW-h/tonne? Is it a show stopper or a just more work for the FEL? What % of the feed might end up in the scats bunker? Yes the p80 being presented to the ball mill could be less, 8-10 mm as you mentioned but I'm looking at this coarse transfer size case.

If feed coarser than the selection function maximum which is at - 2 mm and ball mill is biggest, it is possible to increase circulating load to get an acceptable product P80 for ball mill. Need to check the pumps and hydro-cyclones to increase the circulating load.It is also possible to increase the number of balls of larger diameter in the mill, if the inverter is slightly increased speed.

I also agree with increasing the circulation load can assist to achieve the required p80 but do not forget that high circulating load will also lead to lower than targeted throughputs in terms of milling. The mill will be backing up due to high circulating load. Normally you need to determine the bond work index for the material you will be treating. With proper Bond work index data then your desired p80 and Milling throughputs will be easily achieved. Physical characteristics of the ore to be treated is very important and we often look at cyclone efficiency forgetting the Kilowatt hour required to reduce one short tonne of ore of an infinite size to 80% passing 100 microns.

We had designed a similar Plant to yours using closed-circuit secondary crushing only, to save on costs, and closed-circuit milled with 90mm, accepting a scats "loss"/lower grade reintroduction, as a consequence of preferential grinding. This concentrator was uprated later to series milling when increasing throughput to beyond 2.5x. I am sure that the ball size should improve the selection function peak; however, tertiary crushing/HPGR should give you a much better product than 13mm, still dry screening? Whereas, secondary closed circuit crushing to 16-18mm should be possible?

But it's a wet grinding process. We're making a 3D model of the system and applying value engineering as we go. For instance the secondary cone crusher feed chute, the screen and the discharge chute add 20 m in length (and two trestles) to the conveyor; so we would now cost the screen and additional chutes, conveyor and structural steel against a cone crusher that can handle the volume of the unscreened feed. We're trying to find the lowest capital comminution system for 1 MTPA. Almost regardless of operating costs!

At the top size you are dealing with crushers are much cheaper to run than ball mills. There are other things to consider. That is hard ore. Your ball size needs to be matched to the ore size. If it turns out you need 4 inch (100 mm) balls to handle the scats your mill liners will last a much shorter time than if you need 2 inch (50 mm) balls. In fact with rubber liners, the 4 inch balls will beat the living daylights out of the liners.

We're working with a plant where exactly what you describe for a crusher circuit going into a ball mill with hard ore is going on. If the crusher gaps or liner wear get out of hand scats can be 10 to 20% of new feed. Unless there is some preferential grinding phenomenon going on and you can throw away the scats, that high rate is not to your advantage.

Don't even think about crushing scats from a ball mill. There is so much metal (ball chips) in there the metal detection system is bypassing the crusher all the time. Of course if your ore isn't magnetic, that isn't a problem. Tramp magnets will pick the worst of the metal.

Finally don't forget to watch what kind of crushers you buy. Some are optimized for aggregate and tend to make a coarse product with a minimum of fines and some are optimized for mining and make a finer product with lots of included fines. You want the latter. The manufacturer may not tell you which is which.

I spread-sheeted the Bond new ball size formula and varied F80 and WI. To me hard is WI>18. So it seems the predicted size should be less than 4". For a 16' diameter ball mill do you have an opinion on what new ball size would be half-way liveable in terms of liner life?

A quick check of the Rowland optimum feed size for this ball mill is between 3 and 4 mm; operating either circuit you describe will be at least 15% inefficient compared a more conventional ball mill feed size. The screen doesn't matter in terms of energy efficiency; only more crushing (or a small rod mill) that will improve the energy efficiency. I doubt the opex benefit of energy efficiency will matter given the capex needed. Ball size sounds right; I get 3.5 inch top size.

I note the question is -6 months old, but there have been some really interesting recent comments.I agree that anything with a BWI>18kWh/t is a hard ore, you mention a WI of -16kWh/t - so I am assuming for my comments that the ore is of medium competency. You also mention a crusher P80 (closed) of 13.4mm or a P80 (open) of 15.4mm. My rough calculation with the above (using the open case) agrees with your ~16' (5.3m) mill size, installed power around 2.5MW.

With regards to 3.5"- 4" media, typically for Ball mills of this size (16'), max size of ball used should be around the 50-60mm. I feel that a 4 ball in a 16 Ball Mill is too large and more suited to SAG Milling. You risk damaging liners (rubber, composite or full steel) as well as trommels and feed chutes in the process. To ensure the final grind, extra EGL in the mill may be required as you rightly point out.

If the EGL is not long enough, you will not get the residence time and hence the grind you are after,The hydro-cyclones will have a hard time if set for the desired P80Increased re-circulating loadLarger than normal feed spouts needed to handle the coarser feedHigh wear on trammel screens

Another option may to consider scrapping the tertiary, and perhaps even secondary crushers (depending on Primary crusher CSS) plus screening / scalping capacities) and using HPGR as secondary crushing / primary grinding before feeding to the Ball mill. There are several iron ore flowsheets in operation that run Primary-Secondary-HPGR (dry), then wet for Ball and Regrind milling before going to product and tails.

By the way this project has been stalled due to tenure issues but the challenge (minimum cost of getting into production, even symbolically) is still worthy of consideration. The design can allow for retrofit of closed-circuit crushing, or quaternaries, HPGR, rods or even SABC but these must be paid for by sold Dore. He has suggested that squeezing down the CSS (P80 9 mm) and having a 4 mm trommel where the oversize could be re-processed (I would say parked) might be do-able at WI 16. One assumption in all this is that a ball mill is available which has lots of grinding length, and even if this were not the case a certain pre- or post-ball mill oversize could be parked providing the capital for parking was less than closing the tertiary or some other solution. The other project setting criterion is that the maintenance and operation should be suited to a remote, let's say island, location.

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henan mining machinery and equipment manufacturer - formula for recirculating load of cyclones and mill

henan mining machinery and equipment manufacturer - formula for recirculating load of cyclones and mill

SBM Machinery is a professional material processing designer and supplier in the world, we have excellent research and development group to provide our clients the ...Circulating Load Formula In Ball Mill, ... load for grinding fundamentals. both the density and flow of the cyclone ... ball mill recirculating load ...

Ore beneficiation equipment, sand making equipment, crushing equipment and powder grinding equipment, which are widely used in various industries such as metallurgy, mine, chemistry, building material, coal, refractory and ceramics.

ball mills

ball mills

Ball Mills What Are These Machines and How Do They Work? Short flash video at bottom of page showing batch ball mill grinding in lab. May have to click on browser "Allow Active X blocked content" to play A Ball Mill grinds material by rotating a cylinder with steel grinding balls, causing the balls to fall back into the cylinder and onto the material to be ground. The rotation is usually between 4 to 20 revolutions per minute, depending upon the diameter of the mill. The larger the diameter, the slower the rotation. If the peripheral speed of the mill is too great, it begins to act like a centrifuge and the balls do not fall back, but stay on the perimeter of the mill. The point where the mill becomes a centrifuge is called the "Critical Speed", and ball mills usually operate at 65% to 75% of the critical speed. Ball Mills are generally used to grind material 1/4 inch and finer, down to the particle size of 20 to 75 microns. To achieve a reasonable efficiency with ball mills, they must be operated in a closed system, with oversize material continuously being recirculated back into the mill to be reduced. Various classifiers, such as screens, spiral classifiers, cyclones and air classifiers are used for classifying the discharge from ball mills. This formula calculates the critical speed of any ball mill. Most ball mills operate most efficiently between 65% and 75% of their critical speed. Photo of a 10 Ft diameter by 32 Ft long ball mill in a Cement Plant. Photo of a series of ball mills in a Copper Plant, grinding the ore for flotation. Image of cut away ball mill, showing material flow through typical ball mill. Flash viedo of Jar Drive and Batch Ball Mill grinding ore for testing Return To Crushing Info Page Contact Us Copyright 1994-2012 Mine-Engineer.Com All Rights Reserved

closed circuit ball mill basics revisited - sciencedirect

closed circuit ball mill basics revisited - sciencedirect

Since the early days, there has been a general consensus within the industry and amongst grinding professionals that classification efficiency and circulating load both have a major effect on the efficiency of closed circuit ball mills. However, the effect of each is difficult to quantify in practice as these two parameters are usually interrelated. Based on experience acquired over the years and the investigative work conducted by F.C. Bond, it was established that the optimum circulating load for a closed ball mill cyclone circuit is around 250%. This value is used as guideline for the design of new circuits as well as to assess the performance of existing circuits.

The role of classification in milling appears to have been neglected in the current efforts to reduce the energy consumption of grinding. Two past approaches, experimental and modelling, for quantifying the effects of classification efficiency and circulating load on the capacity of closed ball mill circuits, are revisited and discussed in this paper. Application to the optimisation of existing circuits and design of new circuits is also discussed, with special attention to the development of more energy efficient circuits.

Circulating load and classification efficiency effect on ball mill capacity revisited. Relative capacity model introduced and validated. Relationship between circulating load and classification efficiency verified by industrial data. Existing fine screening technology could increase ball mill circuit capacity 1525%.

energy and cost comparisons of hpgr-based circuits | e & mj

energy and cost comparisons of hpgr-based circuits | e & mj

A comprehensive energy and cost study compared an existing SAG mill-based circuit at the Huckleberry mine with two proposed circuits involving comminution technologies that are associated with energy efficiency: high-pressure grinding rolls (HPGR) and high-speed stirred mills. The specific energy requirements, expressed as kilowatt-hours per metric ton (kWh/mt), for the proposed circuits were determined from pilot-scale HPGR and stirred mill testing conducted at the University of British Columbia (UBC).

Samples and operating data were collected from Huckleberrys copper-molybdenum concentrator to evaluate current mill performance for comparison. To support the base case, the current Huckleberry mill circuit was modeled using JK SimMet software. The main comparison focused on the complete energy requirements for each circuit, including materials handling equipment such as conveyors, screens, feeders and pumps. Capital and operating cost estimates for each of the comminution circuits are also given.

The results showed that the HPGR-ball mill circuit achieved a 21% reduction in energy consumption over the existing SAG-ball mill circuit at the same P80 grind size of 160 mircons (m). At a grind of 80% passing 75 m, the HPGR-stirred mill circuit showed a 34% reduction in energy compared to the base case. The energy reduction for the new flowsheets significantly improved the economics of the Huckleberry comminution process.

IntroductionUp until now, tumbling mills such as AG/SAG mills and ball mills have had a dominant bearing on the design andeconomics of comminution circuits. However, it is commonly agreed thatthe majority of employed comminution processes are energy intensive and energy inefficient, accounting for up to 80% of overall process plant energy consumption and having an efficiency of as lowas 1%. The U.S. Department of Energy reported that there is a potential to reduce energy consumption in the metals industry by up to 61% from current practice to best estimated practical minimum energy consumption; suggestions included the implementation of best practices and the adoption of energy efficient mining and mineral processing technologies such as advanced blasting techniques, HPGR and stirred mills.

The concept of combining an HPGR and a stirred mill in a single flowsheet has been proposed, which was envisioned to be an example of the future in energy conscious comminution processes. The pilot-scale HPGR and high-speed stirred mill testing facility at the UBC Norman B. Keevil Institute of Mining Engineering provided a very unique opportunity to assess the HPGR and/or stirred mill circuits and understand the potential benefits. To examine a combined HPGR and stirred mill circuit, both machines have to be operated outside their currently accepted operating conditions. Studies have demonstrated that an HPGR-stirred mill circuit is technically feasible and showed promising benefits over the traditional stage crushers-ball mill circuit and HPGR-ball mill circuits.

To determine whether the novel HPGR-stirred mill circuit arrangement could achieve energy savings in comparison to conventional SAG mill-based circuits, a pilot-scale study was conducted to compare the energy requirements of the existing circuit at the Huckleberry mine to two alternative circuits: an HPGR-ball mill circuit and a novel HPGR-stirred mill circuit. The study was conducted in collaboration with the Huckleberry mine with support from BC Hydro, Xstrata Technology and Koeppern.

Data representative of three hours of continuous mill operation, directly preceding mill shutdown and sample collection, was analyzed to confirm process stability and subsequently to determine the actual specific energy requirement of Huckleberry process equipment. A sample was collected and analyzed using established comminution laboratory testing methodologies, characterizing the properties of the ore and slurry for modeling and simulation of the circuit.

The sample consisted of a 1,000-kg SAG belt-cut sample, at nominally 100% passing 100 mm, and two buckets of cyclone overflow slurry. The feed material had a moisture content of 3% with a specific gravity (SG) of 2.76. The battery limits for the energy study were established as being the feed to the SAG circuit (coarse ore stockpile) and cyclone overflow of the ball-mill grinding circuit (feed to flotation circuit).

The collected SAG feed sample was prepared for pilot-scale HPGR and stirred mill testing to determine the key operating parameters for flowsheet design and power-based calculations. Ultimately, the simulation and test results allowed for the direct comparison of the energy and costs for three circuits.

The current process configuration at the Huckleberry operation is based on a SAG mill operating with a pebble crusher and ball mills, commonly referred to as an SABC type comminution circuit (See Figure 1). Modeling of the SABC circuit using JKSimMet was carried out using known equipment parameters, operational data and the results of material analyses as inputs. The JKSimMet model was used to confirm the validity of acquired data and to model the effects of modifying certain areas of the comminution flowsheet.

The HPGR-ball mill circuit comprises a reverse-closed secondary crushing circuit prior to a closed HPGR circuit, followed by a reversed-closed ball-mill circuit with cyclones (See Figure 2). The vibrating screen decks were set to an aperture size of 32 and 4 mm for the secondary crushing stage and HPGR screen circuit, respectively. The energy requirements for the secondary crushing stage and the ball-mill grinding stage were determined using the previously fitted JKSimMet model. A number of pilot-scale HPGR tests were carried out to determine the proper operating conditions. Energy values obtained from pilot HPGR testing, laboratory testing and JKSimMet modeling were combined to calculate the specific energy requirement for this circuit.

The novel HPGR-stirred mill circuit is comprised of a reverse-closed secondary crushing circuit prior to an open HPGR circuit, and followed by a second HPGR in closed circuit to generate finer feed for high-speed stirred milling (See Figure 3). The vibrating screen decks were set to aperture sizes of 32 and 0.71 mm for the closed secondary crushing stage and second stage HPGR screening circuit, respectively. The energy requirements of the secondary crusher were determined using the JKSimMet crusher model. A number of pilot-scale HPGR tests were carried out to determine the proper operating conditions for the first stage of HPGR crushing. Recycle tests were performed to simulate the HPGR performance in closed circuit with a screen and to determine the associated specific energy values. Energy readings obtained from pilot HPGR testing, stirred mill testing and JKSimMet modeling were combined to calculate the total specific energy requirement for the proposed novel circuit.

Comparison of All CircuitsThe battery limit for comparison of comminution energy using the proposed circuits was feed from the coarse ore stockpile with an F80 of 66 mm to two product sizes of P80 of 160 m (current target grind for the Huckleberry mine) and P80 of 75 m. For the target grind P80 of 160 m, the existing SABC circuit was compared to an HPGR-ball mill circuit. For this product size, the IsaMill could not be tested because the second stage HPGR product particle size (screen undersize feeding the IsaMill) was only slightly larger than the target P80, such that stirred mill grinding would be impractical. Therefore, to compare all three circuits, the target grind size was selected to be a P80 of 75 m. (See Table 1 and Figure 4)

For grinding to a P80 of 160 m, the HPGR-ball mill circuit required 21% less energy than the SABC circuit. The main savings result from the lower energy required by the HPGR as compared to the SAG mill. However, an additional secondary crusher and conveyer system were required to facilitate the HPGR circuit. The HPGR also produced a coarser product than the SAG mill. Thus, the energy needed for crushing, ball milling and material handling was higher for the HPGR option than the SABC circuit.

When extending the target grind size to a P80 of 75 m, the energy savings of the HPGR-ball mill circuit was only 7%. The novel two-stage HPGR-stirred mill circuit demonstrated a significant reduction in energy of 34%, as compared to the SABC circuit. It must be noted that the energy for any additional equipment, which would be required to disperse the HPGR product prior to screening at 0.71 mm, was not accounted for. However, assuming equipment similar to that of a vertical shaft impacter (VSI) were to be used prior to screening, an overall reduction in energy in excess of 30% would still be expected. In retrospect, the study shows that the application of an HPGR-stirred mill circuit can significantly lower the energy required for mechanical size reduction.

There was a discrepancy in the additional power required to reduce the final product size from 160 to 75 m for SABC circuit and HPGR-ball mill circuit. The HPGR-ball mill circuit when grinding to 75 um compared to 160 um used an extra 4.6 kWh/mt, but for the SABC circuit it only used an extra 1.8 kWh/mt. The reported SABC circuit power for the 160 m grind size was based on actual site data. JKSimMet was used to fit a model to site DCS data. Thus, there was considerable scope to reduce the ball-mill power consumption by using smaller ball-mill grinding media. However, for a final grind of 75 m, the reported power values for both HPGR-ball mill and SABC circuits were based on JKSimMet modeling alone. In these cases, the previously fitted model was optimized through process design changes (media size, transfer size, cyclone parameters etc.) to achieve the 75 m product size while minimizing energy consumption. So the difference in ball-mill power to reduce the final grind size from 160 to 75 m should have been much greater in the case of the SABC circuit than actually reported.

Capital and Operating CostsTo complete the comparison of the process options, capital and operating costs were determined from vendor quotes and installation costs. The costs are deemed to have an accuracy of 50% to a preliminary level of assessment (See Table 2 and 3). For comparison, the costs for SABC circuits grinding to 160 m and 75 m were determined to allow for direct comparison to HPGR-ball mill and HPGR-stirred mill circuits, respectively.

The indirect cost was estimated at 45% of direct capital costs and was considered to be within industry standards for the options considered. A similar approach was applied to estimate the total direct costs. It was found that both HPGR-ball mill and HPGR-stirred mill circuits have higher associated capital costs than the SABC option. Conversely, operating costs for the two proposed circuits are substantially lower, which relates directly to lower energy consumption levels.

The trade-off economics were calculated on the basis of net present value (NPV). A discount rate of 5% and a 15-year mine life were assumed. At a grind of 80% passing 160 m, the HPGR-ball mill circuit shows significant cost advantage over the SABC circuit with a NPV of $33 million and an IRR of 22% (See Table 4). At a product size of 80% passing 75 m, both options have costs advantages over the SABC option, although the HPGR-ball mill circuit had lower overall costs than the two-stage HPGR-stirred mill circuit. In general, a finer grind size would not be selected unless it resulted in significant recovery improvements. Since the copper-molybdenum recovery versus grind size information was not available, such a comparison was not possible for this study. However, in cases where a finer primary grind is needed to achieve high metal recoveries, the two-stage HPGR-stirred mill process demonstrates significant energy savings that would be reflected in the NPV.

DiscussionEvaluation of the proposed circuits showed that combining the two comminution technologies, HPGR and stirred mill, had considerable potential as an energy efficient and economic approach to grinding metallic ores. Overall, the proposed HPGR-based circuits were found to be more energy efficient than the current Huckleberry SABC circuit. Both the higher energy efficiency and elimination of steel grinding media associated with the HPGR-based circuit significantly reduced the determined operating costs.

The effect of ore variability was not evaluated in this study; however, HPGRs are certainly less sensitive to variation in ore hardness when compared to SAG mills. There is also a question of the differences in liberation characteristics from a SAG mill-ball mill operating in closed circuit with a classifying cyclone and the HPGR-stirred mill operating in open circuit. With differences in particle breakage mechanisms as well as mineral particle size distributions (since no cyclone classification is used in stirred mill operation), it would not be surprising to find differences in degree of liberation.

During pilot simulation of the HPGR-stirred mill circuit, both machines were operating outside their respective industry standard conditions. Challenges were primarily associated with the nominated transfer size between the HPGR and stirred mill. For example, nomination of a coarser transfer size necessitated the use of larger stirred-mill grinding media and resulted in a reduction in stirred-mill energy efficiency. Conversely, nomination of a finer cut-point was detrimental to the screening efficiency of HPGR product. Successful development of the HPGR-stirred mill circuit relies on further addressing the efficient separation of HPGR product at a suitable feed size for stirred mill operation. This will likely involve the introduction of an additional piece of material dispersing equipment that would be located prior to the HPGR screens.

Conclusions and RecommendationsThe presented study built on previous related work at the UBC NBK Institute of Mining and showed that an HPGRstirred mill grinding circuit, as an alternative to commonly implemented SABC comminution circuitshas significant potential as an energy efficient alternative. A reduction in energy consumption of 34% was determined to be attainable through implementation of the HPGR-stirred mill circuit when targeting a final P80 grind size of 75 micron. Project economics were also in favor of the proposed circuit and would further improve in regions where energy supply is more expensive than the relatively low energy unit costs used as a basis for this evaluation.

Other advantages identified with the proposed circuit include the resilience of HPGRs and stirred mills to changes in ore hardness. Carrying out further pilot tests using samples taken from different areas of the Huckleberry deposit would allow for this attribute to be quantified in terms of its influence on energy requirements and overall project economics.

Further work is required to improve the classification of HPGR product and to optimize stirred mill parameters for treating coarser feed sizes. The former is particularly challenging when taking into account the detrimental effect of moisture on HPGR performance, thereby necessitating a dry classification process to limit the amount of moisture returned to the HPGR grinding section with oversize particles.

The results of this study clearly show that there is considerable scope for improving the energy efficiency of industry standard comminution grinding circuits. The proposed HPGR-stirred milling has demonstrated significant potential as a means to grinding more efficiently, this attribute being increasingly important as the mining industry is faced with extracting metals from harder and more complex deposits.

C. Wang, O. Mejia and B. Klein are with the University of British Columbia in Vancouver. (Wang can be reached at: [email protected]). S. Nadolski is with Koeppern Machinery Australia in Malaga. J. Drozdiak is with Hatch in Vancouver. This article is adapted from a paper, Energy and Cost Comparisons of HPGR-based Circuits with the SABC Circuit Installed at the Huckleberry Mine, they presented at the 45th Annual Canadian Mineral Processors Operators Conference, which took place in Ottawa, Ontario, January 22-24. To view the entire unedited paper, visit the Resource Center at CEEC: www.ceecthefuture.org/resource-center/.

Canadas Finest Recognized with CEEC MedalThe paper from which this article was derived received the2013 Coalition for Eco-Efficient Comminution (CEEC) Medal. Congratulations to Fisher Wang, Stefan Nadolski, Olav Mejia, Jeff Drozdiak and Bern Klein on their excellent work, said CEEC Chair Elizabeth Lewis-Gray. The business of energy efficiency in grinding and crushing is a global issue and we are delighted that the winners of the prestigious 2013 CEEC Medal authored their outstanding paper as a result of an industry first; collaboration between BC Hydro, University of British Columbia and the mine operator Huckleberry mines. The 2013 Medal recipients are based in North America, specifically in Vancouver, and they join previous CEEC Medal recipients based in Peru, Chile and Australia.

The 2013 Medal was awarded in Vancouver, B.C., before a gathering of industry leaders. More than 60 guests applauded as The Right Honorable Dave Nikolejsin, deputy minister for energy and mines, presented the medals to the winners. Ausenco hosted the event at its Vancouver offices.

The CEEC Medal is a global award intended to recognize and celebrate the contribution of outstanding published papers, articles or case studies profiling beneficial strategies for eco-efficient comminution. In 2013, more than 10 papers were nominated. The Medal review committee is led by CEEC Director Dr. Zeljka Pokrajcic, principal process engineer with WorleyParsons. The committee evaluates all nominated papers on the basis of the papers potential to improve energy efficiency, the ability for the concepts to be readily adapted to operating plants or incorporated in the design of new circuits, that the results are robust and believable, and finally, whether or not the paper communicates its ideas clearly and effectively. The medal committee recommends the annual recipient to the CEEC Board of Directors for its approval.

CEEC is a not-for-profit company funded by sponsorship from the mineral industry itself, whose mission is to accelerate knowledge and technology transfer in the field of energy-efficient comminution (crushing and grinding). CEEC aims to raise awareness of beneficial alternative comminution strategies with the objective of improving earnings, achieving lower processing costs and gaining energy efficiencies in the mining sector.

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cyclone monitoring system improves operations at kucs copperton concentrator | e & mj

cyclone monitoring system improves operations at kucs copperton concentrator | e & mj

Kennecott Utah Coppers (KUC) Copperton Concentrator has four grinding lines each consisting of one semi-autogenous SAG mill feeding two ball mills in a closed circuit with a hydrocyclone battery (See Figure 1). The aim of the ball mill/cyclone circuit is to produce the optimum flotation feed particle size while maintaining grind throughput. The optimal operating point is a tradeoff between throughput, recovery and grinding cost.

Copperton recently focused on the role of the cyclone battery and how to monitor the performance of each individual cyclone. Variation in the performance of the grinding circuit flows through to the cyclones and in many cases the cyclones will report coarse material to the overflow when not operating as designed. Coarse material in the flotation feed reduces the economic performance of the concentrator through lower valuable mineral recovery and in extreme cases, through blocking of the flow path in the flotation cells.

The existing oversize detection system is installed on the combined cyclone battery overflow line. Operationally, troubleshooting the cause of oversize is difficult and time consuming, resulting in considerable disruption to the flotation circuit before the offending cyclone is taken offline. Copperton and CiDRA recently installed a new technology for monitoring individual hydrocyclone overflow lines for coarse material discharge.

The Value of Early Detection The target grind size for flotation feed at Copperton is 32% +150 microns (100 mesh). Cyclone product greater than 150 microns is considered coarse (or oversize). Particles 6 mm (~0.25 inches) or greater can be considered rocks and a large flow of this material can block the internals of the flotation cells, significantly impacting their performance. Figure 2 shows an example of the material that has been flushed from a rougher row after a severe rock event. Some causes of rock events include: SAG mill grate or trommel panel failure; cyclone failure, and degradation or failure of SAG mill pebble removal system.

It is difficult to quantify the economic impact of large rock events due the widespread impacts and flow on effects across areas. However, qualitatively the losses are associated with: rougher row downtime (recovery loss through lower residence time), reduced throughput, equipment damage (flotation cell internals, slurry pumps, etc.), cleanup costs, and decay in flotation cell performance over time (event not large enough to lead to shut down).

Recovery of liberated and middling +150-micron mineral particles in the flotation circuit is significantly lower than -150-micron particles. Based on mineral liberation analyzer (MLA) data, the long term average +150-micron copper recovery is significantly lower compared with recovery of -150-micron particles (See Figure 3). It is also worth noting that the recovery of +150-micron size fraction is considerably more variable than for the -150-micron size fraction.

The value lies in being able to detect a rock event early and optimize cyclone performance to reduce the amount of +150-micron particles in the flotation feed. The new technology aims to provide rock detection and identify cyclone failure and the cyclone(s) producing rocks.

Development of New Monitoring System One approach to confront the previously described grind challenges is to optimize grind system performance at the individual cyclone level, rather than the battery level. This approach has the advantage of allowing a cyclone battery to remain in operation while a performance issue is isolated to one or more individual cyclones. However, this requires a new class of instrumentation, one that monitors the performance of each cyclone in real-time.

One of CiDRA Mineral Processings core competencies is the measurement of acoustic information through the wall of a pipe. This technology is not only well suited for monitoring individual cyclones, but is also non-intrusive. The sensor is attached to the external surface of the pipe, allowing sensors to be installed and maintained without interruption to the process.

CiDRAs acoustic measurement technology, CYCLONEtrac, provides a novel solution for problems such as rocks reporting to the overflow and unreliable feed isolation valve position (limit switch failures). Additionally, the system provides the control room with an indication of the operating mode of each cyclone. The system can differentiate between a cyclone that is off (feed isolation valve shut), operating normally (fines reporting to overflow and coarse to underflow), or abnormal states such as rocks reporting to the overflow.

Prototype hardware was first installed on a cyclone battery at Copperton to record real-time acoustic data directly to a mass storage device. The data was post processed and used to develop the initial CYCLONEtrac algorithms. Then the new systems performance was compared to Coppertons existing oversize material detector installed in the combined overflow of the battery. The detection algorithm was refined and the hardware and software designs were validated.

A full system was then built and installed on all eight cyclone batteries in the grind plant. Figure 4 shows the transmitters and junction box for one battery, the CYCLONEtrac band installed on an overflow pipe, and a full battery instrumented with bands.

The system consists of a sensor and preamplifier attached to the external pipe surface on the overflow of each cyclone. A schematic of the system is shown in Figure 5. The preamplifier outputs are connected to transmitters which perform the first level data processing for each battery. The data accumulated by the transmitters is aggregated by a computer in the Copperton control room. This data is used by the CYCLONEtrac system to determine the operational state of each individual cyclone and of each battery of cyclones. In addition to the real-time display, data is stored for local review and transmitted to the CiDRA data center for the generation of daily and weekly performance and utilization reports. Both the displayed data and the summary reports facilitate smart maintenance scheduling and aid in troubleshooting to reduce the downtime associated with periodic maintenance and repairs.

System Validation Throughout the development of the CYCLONEtrac system, real-time acoustic data for both normal and abnormal operating conditions were recorded, including 15 severe rock events. One of the events occurred on February 3, 2009, and is analyzed here to demonstrate the system response.

Using the defined system threshold parameters the system successfully detects a sustained rock event. This data was then compared to Coppertons oversized particle detection system data from the data historian. The CYCLONEtrac system detected this event as displayed in Figure 6. The blue trace is the sound pressure level (SPL) calculated by CYCLONEtrac. The blue trace shows that the SPL of the acoustic signal for hydrocyclone No. 3 increases more than 6 dB during the event. The overlaid maroon bars represent the number of impacts detected by CYCLONEtrac for each minute of the event (from rocks striking the inner pipe wall), as defined by the algorithm.

This event was isolated to a single cyclone, as shown on a screen shot of the control room interface in Figure 7, and the excursion in SPL and rock count is not detected on any of the other cyclones. The solid black line plotted with SPL is from Coppertons oversize particle monitor (in percentage of sample) from the combined overflow of all 10 cyclones in the battery. Thus an event that would have been generally associated with the whole battery has been isolated by CYCLONEtrac to the single cyclone affected.

The event terminated rapidly (at 800 seconds on the plot) when feed flow stopped to the effected cyclone by closing the isolation valve. The SPL dropped very quickly while the number of rocks and the oversize detector signal roll off more slowly due to slower sample rates.

When rock events occur, the CYCLONEtrac system will alert the control room operator that rocks are reporting to the flotation system rougher banks allowing the operator, or the control system, to correct the problem without shutting down the entire battery.

A short sampling campaign completed in July 2011 was designed to further confirm the performance of the system. A sample basket assembly was attached to the overflow discharge of two cyclones on separate batteries for 24 hours and 48 hours, respectively. During these two sampling periods, the CYCLONEtrac system determined the number impacts on the pipe wall, indicating that rocks were reporting to the overflow of the particular cyclone under test. At the conclusion of the sampling period, the material collected in the basket was removed, dried and analyzed. There was a correlation between the number of rocks (wall impacts) detected and the number of rocks collected. Of interest is that at the point where the rocks were detected, the combined overflow % oversize detector signal also spiked.

The number of rocks large enough to be detected was small, so it is only an assumption that the rocks collected were those causing the impacts. The highlighted region on the left side of Figure 8 shows the corresponding rise in the output of Coppertons oversize particle detector, at about 2:24 on July 7, 2011, and the two wall impacts that were detected by CYCLONEtrac at the same time (assumed to be the two large rocks that were recovered from the sample basket).

Another simple verification test was performed by tapping on the overflow pipe adjacent to the sensor with a metal wrench to simulate wall impacts. While the impacts on the outside of the pipe cannot completely simulate those created by rocks on the inside of the pipe, they do test the capability of the system to respond to a series of sharp impacts. As a result of this test, the CYCLONEtrac system detected a rock event on cyclone No. 3, on the right side of Figure 8, at 10:00 on July 8, 2011. When the rock event was detected, it triggered an alarm in the Copperton control room and initiated notifications to the CYCLONEtrac development team. Note that the oversize particle detector did not respond in this case as there was no change to the distribution of particles reporting to the overflow of the cyclone.

Value Realization The measurement system was developed to support detection of rock events, however the main focus for Copperton was identifying and resolving system and operating issues to eliminate the source of oversize. Since the full installation of the CYCLONEtrac system there has been no significant oversize events to fully realize the value of the CYCLONEtrac. In parallel with the CYCLONEtrac development, beginning in 2009, root cause analyses on some oversize events identified the SAG mill discharge as a problem. As a result Copperton asset maintenance tactics were improved, with a greater focus on the health of:

Operations tactics around oversize events also improved. More attention has been paid to equipment inspections and the response to events has been more immediate due the appreciation at the operating floor level of the significant economic impacts of such events.

The continued focus on the root cause of oversize events and immediate response to early indicators (in the SAG mills) has dramatically reduced large events impacting flotation. Nevertheless, the Copperton Concentrator now has a system that will detect rock events, is non-intrusive to the process and requires limited maintenance.

The CYCLONEtrac system has yet to be formally integrated into Copperton operations, primarily due to the fine tuning of the threshold parameters. The response to a cyclone in alarm state has been for an operator to sample the individual overflow stream for oversize and then sample the remaining cyclones in the battery. If the sample confirms there is oversize, the cyclone is isolated. If no oversize exists, the alarm is acknowledged and normal operation continues. The operating strategy and control response plans for such events require finalization and all operating crews will then be trained on the system.

Importantly, performance data is now available for each individual cyclone rather than the battery as a whole unit operation. The full potential of this data hasnt been fully explored but potential exists for data utilization from the operating mode output to be incorporated into routine cyclone maintenance and replacement frequency.

Dylan Cirulis is a metallurgist working for Kennecott Utah Copper and Jerin Russell is a senior development manager for CiDRA Minerals Processing. This article was adapted from a presentation Dylan Cirulis made at E&MJs 2011 Mineral Processing, which was held during October in Lake Tahoe, Nevada, USA.

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mill circuit cyclones | multotec classification cyclones | multotec

mill circuit cyclones | multotec classification cyclones | multotec

Mill circuit cyclones from Multotecmaximise mill efficiencyandreducetheenergyper ton required. This is achieved via theirunique inlet designand, additionally, the cyclone configuration can be altered to accommodate changing operational conditions.

Improved mill capacity by this cyclone, part of Multotecs range ofclassification cyclones, facilitates increased throughput. The specifically designed cut points are easily achieved with a single cyclone size or distributor mounted cluster cyclone configurations.

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