After the grinding circuit has been brought up to normal operating conditions, the operator must monitor the various process variables and alarms. Most of these variables are monitored in the mill control room, however, the operator is also required to sample and analyse process streams and read local indicators.
The ball mill is susceptible to variations in ore hardness resulting in various grinds at constant throughput, or alternatively, various tonnages at constant grind. The variation in grind is not directly determined. However, a changing cyclone overflow density, at a constant tonnage rate and feed density, would be indicative of changing ore hardness and in that case the tonnage fed to the ball mill should be changed accordingly.
The ore feed rate to the ball mill is controlled by the weightometer located on the mill feed conveyor which can be manually adjusted with in the control room to give a constant weight reading. The signal from the weightometer increases or decreases the belt feeder speed and adjusts the water addition to the ball mill (as a function of the actual weight reading). Both weight control and the proportion of water can be adjusted in the control room.
The water rationing controller must be adjusted programmatically to give the desired ball mill discharge density (normally 60-65% solids). The ball mill discharge density should be checked manually at regular intervals and adjustment made to water ratio controller setpoint to adjust the ball mill discharge density.
The grinding circuit operator must ensure that the ball mill runs properly loaded and gives the correct ore grind. A major practical indication of mill loading is the sound made by the mill. A properly loaded mill will have a deep rhythmic roar, while an under loaded mill will have a metallic rattling type noise and an overloaded mill will be quite silent.
The operator must manually measure the cyclone overflow and underflow densities regularly. An increase in overflowdensity is indicative of softer ore and will soon be accompanied by a lowering of power draw at the mill and a change of sound indicating that the mill is becoming under loaded. To compensate, feed tonnage must be increased. Similarity, a decrease in the cyclone overflow density is indicative of harder ore and this will be accompanied eventually by a coking of the mill. Feed to the ball mill must be reduced.
In the event of an emergency, the mill feed conveyor is shut down individually or by stopping the operating cyclone feed pump. The ball mill must be shut down separately. All equipmentshutdowns are performed locally or from the MCC located in the mill control room.
The simplest grinding circuit consists of a ball or rod mill in closed circuit with a classifier; the flow sheet is shown in Fig. 25 and the actual layout in Fig. 9. This single-stage circuit is chiefly employed for coarse grinding when a product finer than 65 mesh is not required, but it can be adapted for fine grinding by substituting a bowl classifier for one of the straight type so as to enable the W/S ratio of the overflow to be kept below the 4/1 limit usually necessary for flotation. On account of the greater efficiency of the bowl classifier the trend of practice is towards its installation in plants grinding as coarse as 65 mesh.
Single-stage grinding is generally to be recommended for small plants on account of its simplicity. Variations in the size and character of the ore are unavoidable in most plants, but they are, as a rule, very much more noticeable when operations are on a small than when they are on a large scale. Multi-stage grinding as practised in large installations may, therefore, prove impossible to control on a very small scale, and for this reason the simplicity of single-stage grinding is likely to result in a greater overall efficiency than would be obtained with a multi-stage arrangement. When, however, the capacity of a plant approaches or exceeds 1,000 tons per day, two-stage grinding becomes preferable because the effect of normal variations of the ore is less marked and control becomes correspondingly easier.
The usual type of two-stage circuit is shown in Fig. 26, and is one that can be employed for any degree of grinding, although a straight must be substituted for a bowl classifier in the second stage when a 48-mesh product is required. It used to be the practice at one time to omit the first classifier and to pass the feed straight through the primary mill to the secondary circuit, but it was not a good method because either the secondary mill received pieces of ore that were too big or else the primary mill overground a large proportion of the feed. Much better results are obtainable by keeping the coarse ore circulating round the primary circuit, which is set for the efficient grinding of such material, until it is fine enough to be sent to the secondary circuit where the machines are set to grind fine ore more efficiently than coarse. It should be noted that the overflow of the primary classifier is sent to the secondary classifier, not direct to the mill, in order that all material which has been ground fine enough in the primary circuit can be discharged immediately without any chance of its entering and being overground in the secondary ball mill.
So important is it from the point of view of efficiency to get the undersize out of the circuit at the earliest possible moment, whether it is produced in the primary or in the secondary mill, that a special intermediate bowl classifier is often installed between the two stages. Such an arrangement has been found very useful in plants in which improvements in dry crushing practice have resulted in a reduction in the size of the feed to the grinding mills with the result that they have been able to take larger tonnages; the classifiers have then becomeoverloaded, especially in the case of older installations in which both stages were equipped with straight classifiers.
The method of installing a bowl classifier to overcome the difficulty is shown in Fig. 27. This circuit is usually adopted in modern practice, but with a bowl instead of a straight classifier, if necessary, in the closed circuit of the secondary ball mill.
Any increase in the efficiency of classification gives greater economy of power by reducing the amount of ore that is overground, so making a larger proportion of the power required to turn the mill available forgrinding the particles that are still too large. From a theoretical point of view, the ideal method of grinding would consist of a series of ball mills, each in closed circuit with a classifier and each so short that the ore in its passage through the mill would be struck only two or three times by the balls, any undersize produced being removed at once by the classifier ; in this way the chance of a particle being struck again after it had reached the required size would be reduced to a minimum. Practice circuit design approaches the ideal:
Open circuit grinding consists of one or more grinding mills, either parallel or in series, that discharges a finalground product without classification equipment and no return of coarse discharge back to the mill. Some very simplistic examples of open circuit grinding are see below and are made of aRod mill,Ball Mill or aRod mill, ball mill combination.
Not all ores can be ground in an open circuit type ofarrangement. Some conditions which do favor open circuit grinding such assmall reduction ratios,reduction of particles to a coarse, natural grainsize,recirculation of cleaner flotation middlings forregrinding anda non-critical size distribution of the final groundproduct.
Closed circuit grinding consists of one or more mills discharging ground product to classifiers which in turn return the coarse product from the size separation back to the mill for further grinding. In this circuit, grinding efficiency is very dependent upon the size separation effected so care should be exercised in selecting the type and size of classifier used to close the system.
This type of grinding is the most common circuit found in mineral processing facilities, mainly because a lot of ores and product requirements are not suitable for open circuit grinding. Some advantages presented by grinding in closed circuit are that this arrangement usually results in higher mill capacity and lower power consumption per ton of product, it eliminates overgrinding by removing fines early and it avoids coarse material in the final ground product by returning this material to the mill.
Although closed circuit grinding offers many choices for arrangement of the equipment as well as combinations of equipment, some of the more common circuits arerod mill/classifier,Ball mill/Classifier,Rod mill/Ball mill/Classifier and Rod mill/Classifier/Ball mill/Classifier.
The importance of the grinding circuit to overall production in any facility should be obvious by now. Because of the responsibilities assigned to grinding it becomes essential that a grinding mill accepts a certain required tonnage of ore per day while yielding a product that is of a known and controllable particle size. This leads to the conclusion that close control over the grinding circuit is extremely important.
There are many factors which can contribute to fluctuationsin performance of a mill, but some of the most common found inindustrial practice are thechanges in ore taken from different parts of the mine,changes in crusher settings,wear in the crushers,screen damage in the crusher circuit.
These are a few things that operators should look for when changes in mill performance are noticed. Stockpiling of ore ahead of the mill can aid in smoothing out some of the fluctuations although it must be stored in such a manner that no segregation occurs.
The reduction ratio in the grinding section is so much greater than in the crushing plant that labour becomes a relatively small item and the power and steel consumption the largest items of cost. Table 18 gives the average total consumption of power that may be expected in modern ball mill installations of various capacities up to 4,000 tons per day, the figures being based on an average medium-hard ore.
The cost of grinding is more difficult to predict than that of crushing because variations in the hardness and toughness of the ore produce proportionately wider variations in the consumption of power and steel. An approximate guide to grinding costs; they are direct costs and include no overhead charges. Power is assumed to cost 0.075 per kilowatt-hour in the case of the smallest plant and to decrease to a minimum per kilowatt-hour for the largest.
As the tonnage rises up to 1,000 tons per day the costs fall rapidly. In plants of greater capacity, however, they do not decrease in the same proportion with increase of tonnage, because the extra capacity is not obtained by increasing the size of the individual machines but by installing two or more similar units side by side, each of equal efficiency.Reduction of costs then becomes more a matter of organization than of plant design.
As already stated, the power and steel costs are the two largest items, those of labour and supplies being small by comparison ; it is on this account that recent progress has been mainly directed towards reducing the consumption of power and steel by means of greater efficiency in classification and by the use of mills of larger diameter.
The way in which the efficiency of classification has been increased has been described in the paragraph headed Grinding Circuits. An increase in the diameter of a mill gives greater economy in two ways : In the first place, the balls do more effective work in a large than in asmall mill, because, falling from a greater height, they shatter the pieces of ore with greater force ; in the second place, the ratio of the deadweight of the mill to the weight of the ball charge decreases as the diameter increases and thus in a large mill the useless weight to be moved is distributed over a greater weight of useful ball load than in a small mill, with the result that a larger proportion of the total power consumption is available to give the balls the cascading and rolling action necessary to break up the ore.
It is essential for the grinding and flotation sections of a plant to be run continuously. It takes nearly half an hour to clear the circuit of even a small grinding unit preparatory to stopping it, and often an hour is necessary to get the circuit fully loaded after restarting ; most of the power required to keep the machines running during the stopping and starting periods is wasted. Moreover, the operation of the flotation machines is poor during these periods so that much of the power required to keep them running is also wasted. In addition, modern practice aims at the elimination of everything likely to cause fluctuating conditions. For these reasons it is the universal custom to run the grinding and flotation plants for 24 hours per day.
The basic parameters used in ball mill design (power calculations), rod mill or anytumbling millsizing are; material to be ground, characteristics, Bond Work Index, bulk density, specific density, desired mill tonnage capacity DTPH, operating % solids or pulp density, feed size as F80 and maximum chunk size, productsize as P80 and maximum and finally the type of circuit open/closed you are designing for.
In extracting fromNordberg Process Machinery Reference ManualI will also provide 2 Ball Mill Sizing (Design) example done by-hand from tables and charts. Today, much of this mill designing is done by computers, power models and others. These are a good back-to-basics exercises for those wanting to understand what is behind or inside the machines.
W = power consumption expressed in kWh/short to (HPhr/short ton = 1.34 kWh/short ton) Wi = work index, which is a factor relative to the kwh/short ton required to reduce a given material from theoretically infinite size to 80% passing 100 microns P = size in microns of the screen opening which 80% of the product will pass F = size in microns of the screen opening which 80% of the feed will pass
Open circuit grinding to a given surface area requires no more power than closed circuit grinding to the same surface area provided there is no objection to the natural top-size. If top-size must be limited in open circuit, power requirements rise drastically as allowable top-size is reduced and particle size distribution tends toward the finer sizes.
A wet grinding ball mill in closed circuit is to be fed 100 TPH of a material with a work index of 15 and a size distribution of 80% passing inch (6350 microns). The required product size distribution is to be 80% passing 100 mesh (149 microns). In order to determine the power requirement, the steps are as follows:
The ball mill motorpower requirement calculated above as 1400 HP is the power that must be applied at the mill drive in order to grind the tonnage of feed from one size distribution. The following shows how the size or select thematching mill required to draw this power is calculated from known tables the old fashion way.
The value of the angle a varies with the type of discharge, percent of critical speed, and grinding condition. In order to use the preceding equation, it is necessary to have considerable data on existing installations. Therefore, this approach has been simplified as follows:
A = factor for diameter inside shell lining B = factor which includes effect of % loading and mill type C = factor for speed of mill L = length in feet of grinding chamber measured between head liners at shell- to-head junction
Many grinding mill manufacturers specify diameter inside the liners whereas othersare specified per inside shell diameter. (Subtract 6 to obtain diameter inside liners.) Likewise, a similar confusion surrounds the length of a mill. Therefore, when comparing the size of a mill between competitive manufacturers, one should be aware that mill manufacturers do not observe a size convention.
In Example No.1 it was determined that a 1400 HP wet grinding ball mill was required to grind 100 TPH of material with a Bond Work Index of 15 (guess what mineral type it is) from 80% passing inch to 80% passing 100 mesh in closed circuit. What is the size of an overflow discharge ball mill for this application?
This study is conducted with the aim of investigating the efficiency of open and closed-circuit molybdenite ore comminution processes (primary and secondary mill, respectively), through mineralogical study of mills feed and product. For this purpose, particle size distribution, minerals distribution, degree of liberation and interlocking of minerals in mills feed and product were studied. According to the results, chalcopyrite, molybdenite, pyrite and covellite constitute the major part of the mineral composition of open-circuit mill feed. Minerals at the mill product, in the order of abundance include liberated molybdenite particles, liberated chalcopyrite and interlocked chalcopyrite with pyrite, liberated and interlocked pyrite particles, and associated silicate gangues. The d50 values of the feed and product particles of the open-circuit mill are equal to 13.80 and 13.40 microns, respectively. Degree of liberation of molybdenite for the feed and product of this mill is almost the same and is equal to 98.0%. Closed-circuit mill feed includes, in order of is abundance, liberated molybdenite particles in the form of blades and irregular polygonal shapes, liberated and interlocked chalcopyrite, and liberated and interlocked pyrite particles with gangue minerals. Molybdenite particles in the mill product are almost completely liberated, and the degree of liberation values of chalcopyrite and pyrite are 84.40% and 91.40%, respectively. According to particles size distribution of the feed (d50 equal to 25.03 microns) and the product (d50 equal to 24.24 microns) of closed-circuit mill, it can be stated that comminution is not well-operated in closed-circuit mill due to the low solid percentage of closed-circuit mill feed and the inefficiency of hydrocyclone. Examination of Mo, Cu, and Fe grade variations for 10days in both off and on modes of mill shows that closed-circuit mill does not have an impact on comminution process. It can even be concluded that the mill has a destructive effect the flotation process by producing slimes.
Liberation of valuable minerals from associated gangue minerals is an important and fundamental step in separating an ore mineral from gangue during physical or physicochemical separation processes. Crushing and grinding processes are typically used by crushers and mills to liberate minerals, which are energy-intensive processes (especially fine grinding by mill). Meanwhile, ball mills are known for their lowest energy efficiency. The efficiency of ball mills is about 1.0% and, in some cases, less than 1.0% based on energy consumption.
Mineral comminution theories are often based on the relationship between the size of the primary feed particles entering the mill and the energy consumed (Eq.1); in most of these relationships, it has been assumed that the ground material is brittle1. In other words, the grindability indices of the minerals in a deposit are typically used for designing a comminution circuit for an ore. These indices are proposed with the assumption of homogenic breakage and continuity of the ore, regardless of the textural properties of the minerals on a micro scale2.
The main missing factor in most comminution theories is the relationship between ore grinding and its textural and mineralogical characteristics. Each ore in the mine has different geomechanical characteristics for various reasons, such as the effect of faults, the presence of dikes, as well as the type of the deposit in different zones of the mine (such as oxidized, supergene and hypogene zones). Variations in geomechanical characteristics will cause different comminution behaviors for ore during blasting and comminution operations by crushers and mills. In other words, due to the diversity of minerals and the difference in their grindability indices, the feed of processing plants includes a distribution of mineral types and particle sizes. Therefore, the study of mineralogy and the textural characteristics of an ore and its host rock will provide valuable information from the perspective of ore comminution behavior and minerals content and thus the design or optimization of their comminution processes.
Ore textural parameters including hardness, minerals liberation degree, particle size, particle size distribution, minerals abundance, type of minerals, interlocking between minerals and mineralogical structures are important in the issue of ore grindability1,3,4, and can be considered as optimizing parameters for comminution processes. Many researchers1,5,6,7,8 have extensively investigated the impact of textural parameters of ores in mineral processing. It should be noted that the combination of these parameters with the operating conditions of concentration processes, especially comminution, is very complex due to the randomized mechanism of breakage in mills. In general, the process mineralogy of the feed and product of comminution process (mill) will lead to presenting solutions to optimize the operation of the existing circuit and corrective suggestions in the flowsheet. In other words, mineralogy of the comminution process leads to the determination of the optimal conditions of the process by providing practical information on the liberation degree of valuable minerals in each size fraction, particle size distribution and how the minerals are interlocked9.
Molybdenite is a sulfide mineral and is often found in copper-molybdenum porphyry deposits, which is processed as a by-product of copper during the flotation process10. Molybdenite mineral is one of the most stable members of Transition Metal Dichalcogenides (TDMs), which are present in the hexagonal system as a 2H poly as well as a 3R. Figure1 shows a schematic image of the MoS2 layered structure along with the XRD pattern crystallographic plates. Sulfur atoms at higher and lower surfaces surround smaller molybdenum atoms in the form of sandwiches11. Molybdenum and sulfur atoms inside the layers are bonded together by strong covalent bonds, but the successive layers of sulfur atoms are joined bonded by weak Van der Waals bonds12.
Due to the crystalline structure mentioned, molybdenite has anisotropic properties, which has led to its different behavior in different faces14, including the fact that the anisotropic property leads to the preferred orientation of molybdenite mineral during grinding; and as the particle size decreases, this orientation increases. In this way, molybdenite is broken under grinding at two different surfaces. Surfaces created by breaking SS bonds (non-polar surfaces) and surfaces resulting from the breakage of strong Mo-S bonds (polar edges)15. It is worth mentioning that in the structure of molybdenite, the bond between SS and Mo-Mo is of Van der Waals type and SMoS is of covalent type. Due to the fact that Van der Waals forces/bounds are relatively weak compared to covalent bonds, it is more likely to break at the edges. On the other hand, the behavior of molybdenite in other concentration processes, such as flotation, is greatly influenced by how it breaks and its liberation degree during milling. The natural floatability of molybdenite is related to its textural characteristics such as flatness, roundness of particles, longitudinal elongation ratio and smooth surfaces. Due to the preferred cleavages along weak SS and MoMo bonds during the grinding process, plate-like fragments are produced from larger particles. Flat and long particles cause poor performance of particle-bubble attachment and thus reduce recovery. Therefore, molybdenite flotation behavior is the result of a combination of the property of natural floatability and particle morphology.
The study of the flotation of copper and molybdenite ores indicates that the recovery of molybdenite and copper flotation is reduced in the coarse, fine and very fine size ranges of these minerals. The highest recovery of molybdenite and copper occurs at sizes 2755 microns16, therefore the optimal grinding of these minerals is of great importance. Due to the anisotropic behavior of molybdenite and its association with other sulfide minerals (and other associated gangue minerals), performing process mineralogy studies can lead to results for proper design or optimization of its comminution circuit. In the present study, the type and behavior of copper sulfide, molybdenite and associated gangue minerals, especially pyrite, have been identified through a process mineralogy approach toward molybdenite comminution circuit (Sungun copper-molybdenum processing complex located in northwestern of Iran). For this purpose, mineralogical studies have been performed on mill feed and product. As a result of these studies, the liberation degree and the particle size, distribution and the interlocking mode of the minerals have been determined. Analysing and combining this information with the operating conditions of the plant led to solutions for optimizing the current comminution circuit. In other words, according to the mineralogy of feed and product of mills, the most optimal operating conditions are determined and implemented in order to improve the efficiency of the circuit.
The present study was performed on the grinding efficiency of molybdenum flotation circuit mills of Sungun coppermolybdenum processing complex. Sungun copper mine and complex with geographical coordinates of 46 43 east and 38 42 north is located in northwest of Iran. In the molybdenum processing plant of Sungun complex (flowsheet is shown in Fig.2) uses two ball mills to perform grinding operations. The primary ball mill operates in an open circuit (after the middle thickener) and the secondary ball mill operates in a closed circuit with a hydrocyclone. The underflow of the middle thickener with a solid percentage of about 55.0% enters the primary ball mill, and the mill product enters the cleaner 3 flotation cells after dilution. The length and diameter of the ball mill used in this department are 2.44 and 1.52m, respectively, which has a slurry mass capacity of 4.13 t/h. Cleaner 4 concentrate is introduced into hydrocyclone clusters (two hydrocyclone clusters, each of which consists of 3, 6-in. cyclone devices) and is separated into a fine overflow fraction with particles size smaller than 38.0 microns and an overflow fraction with particles size larger than 38.0 microns. The overflow of each cluster is transferred to 58 cleaner cells and its underflow enters a regrinding ball mill, which is in a closed circuit with these hydrocyclones. The goal of the regrinding stage is to achieve the maximum liberation degree of molybdenite and copper minerals and their liberation from each other. The length and diameter of the closed-circuit ball mill are 1.83 and 1.22m, respectively, which has a slurry mass capacity of 3.6 t/h and a solid percentage of 5060% (designed for the plant).
In order to study the mineralogy of milling process in the molybdenum processing circuit, samples of feed (middle thickener underflow) and product of open-circuit ball mill, as well as overflow and underflow of hydrocyclone (feed) and product of closed-circuit ball mill were prepared. Sampling points are marked in red in Fig.2. It is worth mentioning that 3 samples were collected using a sampling spoon from each location and at 30-min intervals, (to investigate the effect of plant feed variations). After filtering and drying, 30g of the sample was prepared using a riffle sample splitter to check size distribution, preparation of polished sections and performing microscopic studies.
In order to perform the process mineralogy studies on the samples, optical microscopic studies were performed on polished sections after ehaviour the size of the feed and product particles using Laser Particle Size Analyzer (SLS: mastersizer 2000/Malvern Panalytical technology). The results of particle size analysis of feed and product samples from grinding circuits are shown in Fig.3, and comparison of their d10, d50 and d90 values are performed in Table 1. According to the results, grinding by open-circuit ball mill has caused the particles size to decrease from 90.0% smaller than 45.0 microns to about 99.0% smaller the 45.0 microns. Grinding by closed-circuit ball mill has also reduced d90 value of particles from 67.96 (mill feed) to 65.13 microns (mill product).
Microscopic study of polished sections is the most common method of studying the mineralogical properties and textural association between minerals in mineral samples. In order to investigate the grinding ehaviour of the minerals in the molybdenum processing circuit, microscopic studies were conducted on polished sections prepared from the collected samples. Microscopic studies were performed using the Leitz polarizing microscope of model SM-LUX-POL equipped with a digital imaging camera at the college of Mining Engineering, University of Tehran. Based on the results of mineralogical studies, open-circuit mill feed contains chalcopyrite, molybdenite, pyrite and covellite. Besides, molybdenite, chalcopyrite, and pyrite are the major minerals that make up the feed of closed-circuit mill (hydrocyclone underflow). Table 2 shows the important physical/chemical properties of various minerals in the feed of open and closed-circuit mill of the molybdenite processing circuit.
The feed of the primary ball mill or the open-circuit mill is the underflow pulp from the middle thickener (Fig.4). According to Fig.2 circuit, the middle thickener with a diameter of 12m and free settling mechanism (in the molybdenum plant of the Sungun coppermolybdenum Complex) is located after cleaner 2 and before the open-circuit mill. The feed pulp to this thickener has a solid percentage of 13.87, which after settling, the underflow is discharged with a solid percentage of about 60.0% and the overflow weight percent is 0.04%. Based on the grade analysis performed on the underflow of the thickener or open-circuit mill feed, the grade values of Mo, Cu and Fe elements are 23.71, 17.57 and 14.84%, respectively. Due to the 23.71% value of molybdenum grade, this product cannot be supplied as a final concentrate and it is necessary to perform more processing stages (cleaner flotation stages). Therefore, the purpose of grinding at this stage is to achieve more liberation of copper minerals from molybdenite.
According to the PSD diagram shown in Fig.5 as well as Table 1, the grinding process in open-circuit mill produces about 70.0% of the fine product with particle size smaller than 20.0 microns; of this amount, 40.0% is smaller than 10.0 microns in size. Particles with a size smaller than 7.0 microns also have a significant volume and account for approximately 25.0% of the mill product particles. Based on the results, it can be concluded that the highest amount of grinding occurred for particles in the size ranges of d75-d25 of feed. Grinding also occurred for feed particles smaller than d25 (7.0 microns), but grinding did not have a desirable result for particles larger than 23.0 microns. Given that the product of open-circuit mill is the feed to cleaner 3 flotation cells, the size distribution of the mill product particles (in other words, the amount of grinding in the mill) is of great importance. Since with increasing the amount of particles smaller than 10.0 microns, flotation recovery of molybdenite gradually decreases due to reduced probability of collision and attachment to air bubbles 17.
According to optical reflected light microscopic studies, chalcopyrite, molybdenite, pyrite, and covellite (Table 2) make up the major part of the composition of the middle thickener underflow or open-circuit mill feed. Liberation studies of molybdenite in the feed of open-circuit mill (Fig.6A) indicate that this mineral has achieved a proper liberation degree (about 98.0%). Therefore, at this stage, grinding leads to more fine production of molybdenite particles and does not cause a significant change in their liberation degree. The study of PSD of the mill product also confirms this; In other words, at this stage, the particles of molybdenite and copper sulfides have become finer and have turned into slimes.
(A) Molybdenite liberated particle in feed, (B) Distribution of minerals in the product, and (C) Molybdenite liberated particles in the product of open-circuit ball mill (Mo molybdenite, Cpy chalcopyrite, Py pyrite).
As can be seen in Fig.6B, the minerals present in the product of open-circuit mill, in the order of abundance, include molybdenite, which is mostly free and has become fine after grinding in mill, liberated chalcopyrite particles (about 32.0%) and interlocked with pyrite and other associated gangue minerals, and liberated and interlocked particles of pyrite. Liberation studies for chalcopyrite and pyrite in the product of mill shows 85.0% and 85.40% values of liberation degree for these minerals, respectively. It can be stated that these two minerals have the same liberation degree. On the other hand, there was no significant interlocking between copper sulfide minerals and molybdenite (Fig.6C). Coarse particles of molybdenite are observable in some cases, but in general the molybdenite particles are ground to very fine size ranges (slime range).
Due to its anisotropic properties, molybdenite behaves differently from other sulfide minerals. SEMTEM images (Fig.7) show how MoS2 breaks (cleaves) in layers. Because in molybdenite with hexagonal structure, the SMoS layers are connected with the covalent bond by the Van der Waals forces. Once molybdenite is ground, its cleavage occurs more easily within the weak Van der Waals forces. Hence, the outer layers of the mineral surface are removed under the influence of shear forces; While the compressive forces in the grinding environment affect the mineral edges and cause breakage in the direction of the edges. The above-mentioned grinding processes, which reduce the thickness of the mineral, occur in the early stages of grinding. As the grinding time increases and in the final stages of grinding, the particle size of the produced layers decreases at a slower rate when compared to the early stages of grinding13. As the size of the molybdenite particles decreases, the surface-to-edge ratio decreases, resulting in an increase in its hydrophilic properties, which reduces its floatability.
The cleaner 4 concentrate with a molybdenum grade of 39.34% and a copper and iron grade of 4.94% and 10.34%, respectively, must be re-floated in order to achieve a higher grade (Fig.2). In this regard, this concentrate is introduced into the hydrocyclone with a separation limit of 38.0 microns, and the underflow is rejected to the closed-circuit ball mill for regrinding (Fig.8). The goal of the secondary grinding stage is to achieve the maximum liberation degree of molybdenite and copper minerals from each other. Examination of the hydrocyclone underflow size distribution (ball mill feed) and mill product (Fig.9 and Table 1) shows that grinding did not have much effect on reducing particle size. In general, the highest grinding occurred in the size range of feed d25d50 values.
Microscopic studies have been performed on mill feed (hydrocyclone underflow) and product. Figure10A,B shows a picture of the minerals distribution in the studied samples. According to studies, the composition of hydrocyclone underflow (mill feed), in order of abundance, includes liberated molybdenite in the form of blades and polygonal fragments (Fig.11), liberated and interlocked chalcopyrite with gangue minerals and liberated and interlocked pyrite particles. The liberation degree of chalcopyrite and pyrite in mill product are 84.40% and 91.40%, respectively. It is worth mentioning that despite the performed grinding, interlocking between molybdenite and chalcopyrite have rarely been observed. In general, according to mineralogical studies and PSD of feed and product of closed-circuit ball mill, the minerals in feed and product are almost similar in terms of particle liberation degree and only particle size has become finer.
As mentioned, the cut size of hydrocyclone in closed circuit with the ball mill is 38.0 microns, however, according to the microscopic images of the hydrocyclone underflow sample (Fig.12), small particles are also observed in this sample, which is due to the inefficiency of the hydrocyclone. Figure13 shows the particle size distribution diagram for hydrocyclone feed, overflow, and underflow, which indicates poor classification performance of the hydrocyclone in the circuit, in the separation of fine particles. According to PSD diagram, d90 value of feed, overflow and underflow of hydrocyclone are 62.0, 55.0 and 68.0 microns, respectively. On the other hand, by measuring the solid percentage of feed (12.0%) and underflow (15.0%) of hydrocyclone, used with the aim of dewatering and particle size control, it can be concluded that this device did not have a good dewatering performance 18. Given the above, the low solid percentage of feed and the abundance of fine particles in the feed can be considered factors responsible for the poor performance of the mill.
In order to more accurately examine the performance of closed-circuit mill, the results of the final molybdenum concentrate analysis were studied in two modes of mill-on and mill-off in a 10-day period (each day including three 8-h shifts). Figure14 shows diagrams for grade analysis of molybdenum, copper, and iron in two modes of mill-on and mill-off for closed-circuit ball mill. According to the figure, the plant circuit is operating in optimal conditions when the mill is switched off. As the circuit became out of optimal conditions, the closed-circuit mill switched on, but the start-up of the closed-circuit mill did not improve the conditions of the processing circuit. It is worth mentioning that the optimal condition means the grade of molybdenum, copper and iron in molybdenum concentrate is more than 50.0% and less than 1.0% and 3.0%, respectively (according to the standards of commercial markets). According to the results, the average grade of molybdenum, copper and iron in molybdenum concentrate in mill-on (non-optimal) mode is 51.21%, 1.32% and 3.95%, respectively. While the average values for Mo, Cu and Fe grade in mill-off mode are 53.83%, 0.71% and 2.04%, respectively. As can be seen, in mill-off mode and optimal conditions, grade standards for molybdenum, copper and iron are available in molybdenum concentrate. However, in non-optimal conditions, starting of the mill has no effect on the optimization of these values, and this indicates the inefficiency of the mill.
Grade analysis of the final molybdenum concentrate for elements (A) molybdenum, (B) copper and (C) iron in the off and on mode of closed-circuit ball mill in a period of 10days (every day includes 3, 8-h shifts).
It is important to know the type and properties of the minerals in the ore being processed in order to design and optimize the circuit of a processing plant. In this study the efficiency of grinding was investigated by studying the mineralogical properties of feed and product streams to the grinding circuits in the molybdenum processing plant. Analysis of particle size distribution for open and closed-circuit ball mills feed and product showed that d90 value of feed and product of open-circuit mill is 43.83 and 42.33 microns, respectively, and d90 value of feed and product of closed-circuit mill were 67.95 and 65.13 microns. The performed liberation study also shows that in the feed and product of the open-circuit mill, the liberation degree of molybdenite is almost the same and about 98.0%. Therefore, in the milling stage, the molybdenite particles are only ground and there is no significant change in their liberation degree. On the other hand, because there is no controlling equipment for particle size in open-circuit mill, fine materials turn into slimes and due to the slime coating, entrainment and less efficient collision of the particles with the air bubble, the flotation rate and the grade is reduced. With excessive grinding of materials in the mill, the surface-to-edge ratio, especially in the case of molybdenite, is reduced, and due to reduced hydrophobicity and floatability, fine molybdenite particles are introduced to tailing product or copper concentrate. In the case of closed-circuit mill, the minerals present in the feed and product of the mill are almost identical in terms of particle liberation degree, and only the particle size gets finer. An examination of the molybdenum, copper and iron grade changes over a 10-day period for both mill on and off modes of closed-circuit mill showed that in the mill-off mode, the plant circuit is in optimal conditions (molybdenum, copper and iron grade in the molybdenum concentrate were more than 50.0% and less than 1.0% and 3.0%, respectively), but as the circuit gets out of optimal condition, the start-up of closed-circuit mill has not had an effect on improving the circuit and creating optimal conditions.
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Bahrami, A., Abdollahi, M., Mirmohammadi, M. et al. A process mineralogy approach to study the efficiency of milling of molybdenite circuit processing. Sci Rep 10, 21211 (2020). https://doi.org/10.1038/s41598-020-78337-8
A specific energy-based ball mill model for continuous operation.The Whiten classification efficiency equation used for the ball mill discharge function.Ore-specific and size-dependent breakage function incorporated in the model.The model was tested with a number of full scale ball mill operation survey data.
A specific energy-based size reduction model for batch grinding ball mills was reported in a previous paper (Shi and Xie, 2015). A discharge function modified from the Whiten classification efficiency equation has been incorporated in the size reduction model to extend its applications from batch grinding to continuous operation. Five sets of the industrial ball milling survey data were used to validate the ball mill model. The data were acquired from four full scale ball mills covering primary and secondary grinding duties in a gold concentrator and a PGM concentrator. In all cases, the model fits the ball mill operational data well.
Features of the specific energy-based ball mill model include the use of an ore-specific and size-dependent breakage function, whose parameters are independently measured with a fine particle breakage characterisation device, the JKFBC. This allows simulations of the effect on ground product size distribution of changing ore breakage characteristics. The model utilises separate selection function and discharge function, which permits the investigation of the influences of mill operational conditions on grinding performance.