The final fineness of the product mainly depends on the number of times the ore particles pass through the grinder. The longer the grinding, the smaller the particle size. Separate crushing and grinding steps are necessary, the ball mill can only receive the broken ore particle, and then grind to the grinding fineness required for flotation.
In order to separate the concentrate from the ore, the ore should be ground fine enough to release the target mineral from the non-mineral grains. The degree of grinding required for this depends on the size of the mineral particles in the ore. A laboratory-scale flotation test is usually required on materials of different particle sizes to determine the grinding particle size required to release the target minerals.The fineness of the ore particles produced by grinding is crucial to recover the minerals by flotation. The most common grinding machines are semi-automatic (SAG) and automatic (AG) mills and ball mills.
Determining an optimal grinding size can maximize the recovery of target minerals in the subsequent flotation process.The grinding size is too large, and some ore particles and non-ore particles cannot be separated, thus preventing their flotation. If the particle size is too fine, the bubbles that rise during the flotation will push the very fine ore-containing particles away, preventing them from contacting the bubbles, thereby reducing their ability to be recovered into the concentrate.In addition, extremely fine rock and iron sulfide particles may agglomerate with extremely fine sulfide ore particles, preventing the ore particles from floating.
According to the test, the particles usually need to be ground to a diameter of about 100 mm to release minerals from each other. When the particles are less than about 10 mm, this is not conducive to the flotation effect.Grinding operations are very power-hungry, which is another reason to avoid excessive grinding.
The crushed products are ground in SAG or AG mills. The self-grinding machine can grind ore without grinding media such as iron ball, or steel rod, as long as the hardness of the ore is sufficient for the rolling ore to grind by itself.A large vibrating screen is used to sieve the ground products to separate the oversized particles. A small cone crusher to recover the oversized material, and then sent them return to the SAG or AG mill for re-grinding. The correct size material is sent to the ball mill for final grinding.
The ball mill is the fine grinding machine connect the SAG or AG mill and flotation machine. Ball mills produce fine particles with a uniform size for flotation, its grinding medias commonly are steel ball. The ball mill rolls grinding media together with the ore, as the ore grinds, these balls initially 5-10 cm in diameter but gradually wear out.Grinding is always carried out under wet conditions, with about 70% solid mixture in water.This procedure maximizes ore production and minimizes power consumption.
A) Total Apparent Volumetric Charge Filling including balls and excess slurry on top of the ball charge, plus the interstitial voids in between the balls expressed as a percentage of the net internal mill volume (inside liners).
B) Overflow Discharge Mills operating at low ball fillings slurry may accumulate on top of the ball charge; causing, the Total Charge Filling Level to be higher than the Ball Filling Level. Grate Discharge mills will not face this issue.
C) This value represents the Volumetric Fractional Filling of the Voids in between the balls by the retained slurry in the mill charge. As defined, this value should never exceed 100%, but in some cases particularly in Grate Discharge Mills it could be lower than 100%. Note that this interstitial slurry does not include the overfilling slurry derived from the difference between the Charge and Ball Filling.
D) Represents the so-called Dynamic Angle of Repose (or Lift Angle) adopted during steady operation by the top surface of the mill charge (the kidney) with respect to the horizontal. A reasonable default value for this angle is 32, but may be easily tuned to specific applications against any available actual power data.
The first step in mill design is to determine the power needed to produce the desired grind in the chosen ore. The most used equation, for this purpose, is the empirical Bond equation (Bond, 1960, 1961; Rowland and Kjos, 1978).
In this equation, E is the specific energy required for the grind, and F80 and P80 are the sizes in micrometers that 80% of the weight passes of the mill feed and product respectively. The parameter Wi, known as the work index of the ore, is obtained from batch bench tests first devised by Bond (1961). The power calculated on using equation 1, (Bond, 1961; Rowland and Kjos, 1978), relates to:
1) Rod milling a rod mill with a diameter of 2.44 meters, inside new liners, grinding wet in open circuit. 2) Ball milling a ball mill with a diameter of 2.44 meters, inside new liners, grinding wet in open circuit.
When the grinding conditions differ from these specified conditions, efficiency factors (Rowland and Kjos, 1978) have to be used in conjunction with equation 1. In general, therefore, the required mill power is calculated using the following equation
where n is the number of efficiency factors, EFi, used and fo is the feed rate of new ore to the mill. The power calculated from equation 2 can be looked up in published tables (Rowland and Kjos, 1978) and the correct mill size and type can be selected.
The philosophy in the development of the MRRC grinding simulation package was to build interactive software that could be used as an inexpensive means of providing a semi-quantitative check on a grinding mill design. In addition the software is designed to slot in to a general mineral processing package now undergoing development at the MRRC.
In this article, alternative forms of optimizing the milling efficiency of a laboratory scale ball mill by varying the grinding media size distribution and the feed material particle size distribution were investigated. Silica ore was used as the test material. The experimental parameters that were kept constant in this investigation was the grinding media filling, powder filling and the mill rotational speed. The data obtained from these batch tests was then analyzed using a model free technique called the Attainable Region method. This analysis technique showed that the required product fineness is a function of grinding media and feed material size distributions. It was also observed from the experimental results that in order to increase the milling efficiency of a ball mill, towards optimum production of material in the desired size class, there is a need to correlate the ball size and the feed size distributions.
Although basic porphyry copper flotation and metallurgy has remained virtually the same for many years, the processing equipment as well as design of the mills has continually been improved to increase production while reducing operating and maintenance costs. Also, considerable attention is paid to automatic sensing devices and automatic controls in order to assure maximum metallurgy and production at all times. For simplicity in this study most of these controls are not shown.Many of the porphyry copper deposits contain molybdenite and some also contain lead and zinc minerals.
Even though these minerals occur in relatively small amounts they can often be economically recovered as by-products for the expense of mining, crushing, and grinding is absorbed in recovery of the copper.
Because the copper in this type of ore usually assays only plus or minus 1% copper, the porphyry copper operations must be relatively large in order to be commercial. The flowsheet in this study illustrates a typical 3,000 ton per day operation. In general most operations of this type have two or more parallel grinding and flotation circuits. For additional capacity, additional parallel circuits are installed.
The crushing section consists of two or three crushing stages with the second or third stages in either closed or open circuit with vibrating screens. Generally, size of the primary crusher is not determined by capacity but by the basic size of the mine run rock. The mine-run ore is normally relatively large as most of the porphyry mines are open pit.The crushing section illustrated is designed to handle the full tonnage in approximately 8 to 16 hours thus having reserve capacity in case of expansion.
Many mills store not only the coarse ore but also the fine ore in open stockpiles using ore as the side walls and drawing the live ore from the center. During prolonged periods of crusher maintenance the ore walls can be bulldozed over the ore feeders to provide an uninterrupted supply of ore for milling.
As it is shown in this study the or 1 crushed ore is fed to a rod mill operating in open circuit and discharging a product approximately minus 14-mesh. The discharge from this primary rod mill is equally distributed to two ball mills which are in closed circuit with SRL Rubber Lined Pumps and two or more cyclone classifiers. The rod mill and two ball mills are approximately the same size for simplified maintenance.
Porphyry copper ores, usually medium to medium hard, require grinding to about 65-mesh to economically liberate the copper minerals from the gangue. Although a clean rougher tailing can often be achieved at 65-mesh the copper mineral is not liberated sufficiently to make a high grade copper concentrate, thus some form of regrinding is necessary on the rougher flotation copper concentrate. It is not unusual to grind the rougher flotation concentrate to minus 200-mesh for more complete liberation of mineral from the gangue.
The cyclone overflow from each ball mill goes to a Pulp Distributor which distributes the pulp to two or more parallel banks of Flotation Cells. These distributors are designed so that one or more flotation banks can be shut down for maintenance or inspection and still maintain equal distribution of feed to the remaining banks.
In some cases it is beneficial to have conditioning before flotation, but this varies from one operation to another and it is not shown in this flowsheet. Ten or more Free-Flow Flotation Cells are used per bank and these cells are divided into groups of four or six cells with an intermediate step-down weir between groups. Free-Flow Flotation Cells are specified, as metallurgy is extremely good while both maintenance and operating expenses are traditionally low. One or more Free-Flow Mechanisms can be stopped for inspection or even replaced for maintenance without shutting down the bank of cells.
The concentrates from rougher flotation cells are sent directly to regrind. Often the grind is 200-mesh. After regrind is flotation cleaning. In some cases the concentrate from the first three or four rougher flotation cells can be sent directly to cleaning without regrinding.
After the rougher flotation concentrate is reground it is cleaned twice in additional Free-Flow Flotation Machines with the recleaned concentrate going to final concentrate filtration or, as the metallurgy dictates, to a copper-moly separation circuit.
The thickening and filtering is similar to other milling operations, however, as the porphyry copper installations are often in arid areas, the mill tailing is usually sent to a large thickener for water reclamation and solids go to the tailings dam.
Automatic controls are usually provided throughout modern plants to measure and control pulp flow, pH and density at various points in the circuit. Feed and density controls are relatively common and the newer installations are using automatic pulp level controls on flotation machines and pump sumps. Automation is also being applied to the crushing systems.
The use of continuous on stream X-ray analysis for almost instantaneous metallurgical results is not shown in thus study but warrants careful study for both new and existing mills. Automatic sampling of all principal pulp flows are essential for reliable control.
The flowsheet in this study illustrates the modern approach to porphyry copper treatment throughout the industry. Each plant will through necessity have somewhat different arrangements or methods for accomplishing the same thing and reliable ore test data are used in most every case to plan the flowsheet and design the mill.
In most plants engaged in the flotation of ores containing copper-bearing sulphide minerals with or without pyrite, pine oil is employed as a frother with one of the xanthates or aerofloat reagents or a combination of two or more of them as the promoter. Lime is nearly always used for maintaining the alkalinity of the circuit and depressing any pyrite present. The reagent consumption is normally within the following limits
While good results are often obtained with ethyl xanthate alone as a promoter, the addition of a small quantity of one of the higher xanthates is frequently found to improve the recovery of those minerals that are not readily floated by the lower xanthate, especially those that are tarnished or oxidized, but since the action of a higher xanthate is, as a rule, more powerful than that of the ethyl compound, it is usually best to add no more of the former reagent than is necessary to bring up the less readily floatable minerals, controlling flotation with the less powerful and more selective lower xanthate. Better results are obtained with some ores by replacing the higher xanthate with one of the dithiophosphates, flotation being controlled, as before, with ethyl xanthate. Sometimes a dithiophosphate can be effectively used without the xanthate, although the dual promotion method is more common. A rule of thumb system for the selection of these reagents cannot be laid down as the character of the minerals differs so widely in different ores ; the best combination can only be found by experiment.When aerofloat is employed alone as the promoter, the reagent mixture is somewhat different from that given above. A reliable average consumption is difficult to determine as the plants working on these lines are few in number, but the following is what would normally be expected.If this combination of reagents gives results equal to those obtainable with a xanthate mixture, its employment has these advantages over the latter method: The control of flotation is not so delicate as with xanthates, it has less tendency to bring up pyrite, and, if selectivity is not required, the circuit may be neutral or only slightly alkaline.
When the ore is free from pyrite, the function of the lime, whatever the reagent mixture, is to precipitate dissolved salts and to maintain the alkalinity of the pulp at the value which has been found to givethe best results ; soda ash is seldom employed for this purpose. When pyrite is present, lime performs the additional function of a depressor, the amount used being balanced against that of the promoterthat is, no more lime should be added than is required to prevent the bulk of the pyrite from floating, as any excess tends to depress the copper minerals, and no more of the promoter should be employed than is needed to give a profitable recovery of the valuable minerals in a concentrate of the desired grade, since any excess tends to bring up pyrite. In many cases a more effective method of depressing pyrite is to add a small quantity of sodium cyanidee.g., 0.05-0.10 lb. per tonin conjunction with lime, less of the latter reagent then being necessary than if it were used alone.
It is not often that a conditioning tank has to be installed ahead of the flotation section in the treatment of sulphide copper ores, as the grinding circuit usually provides suitable points for the introduction of the reagents. The normal practice is to put lime into the primary ball mills and to add xanthates at the last possible moment before flotation, while aerofloat and di-thio-phosphates are preferably introduced at some point in the grinding circuit, since they generally need an appreciable time of contact as compared with xanthates. There is no special place for the addition of pine oil, but care should be taken if it is put into the primary ball mills, as a slight excess may cause an undue amount of froth to form in the classifiers.
In a plant where the primary slime is by-passed round the grinding circuit, it is necessary to ensure that this portion of the pulp receives its correct proportion of and contact time with the reagents.
As regards flotation installations, the present tendency is to employ machines of the air-lift or Callow-Maclntosh rather than of the subaeration type. While two stages of cleaning (circuits 10 and 11) are sometimes essential to the production of a clean final concentrate, circuits 8 and 9 comprising a single stage of cleaning are probably the most widely used. Occasionally the primary machines can be run as rougher-cleaner cells (circuit No. 5), particularly when they are of the air-lift or subaeration type. This method, however, is not often employed, although its use is more common in the flotation of copper sulphide minerals than of any other class of ore ; a stage of cleaning is preferable as providing greater lattitude of control.
Two variations of normal procedure are worth notice. In one or two plants employing two-stage grinding, improved results have been obtained by separating the slime from the primary ball mill circuit and sending it direct to a special flotation section. This method is useful when the feed to the flotation plant contains an appreciable quantity of fines, which, due generally to oxidation through exposure, require different treatment from the unweathered part of the ore. Such fines are usuallyfriable and can be separated as slime from the primary grinding circuit without the inclusion of an undue proportion of unoxidized material, the bulk of which thus passes to the secondary grinding circuit and thence to its own division of the flotation plant.
The second variation consists of grinding the rougher concentrate before cleaning. The method is applicable to an ore in which the copper- bearing minerals are so intimately associated with pyrite that very fine grinding is necessary to liberate them completely. It is often possible, after grinding such an ore to a comparatively coarse mesh, to make a profitable recovery of the copper in a low-grade concentrate which does not represent too large a proportion, say 30% or less, of the total weightof the feed. The concentrate can then be reground and refloated with the production of a high-grade copper concentrate together with a low- grade pyritic tailing suitable for return to the roughing circuit. This method is likely to be less costly than one involving the fine grinding of the whole ore. No standard system can be given for handling the various products as their disposal depends so much on the occurrence of the minerals and the efficiency of the regrinding operations, but a typical flow sheet is illustrated in circuit No. 12 (Fig. 60). It is diagrammatic to the extent that the thickener and regrinding unit may receive its feed from several roughing machines and deliver its discharge to a number of cleaning cells. It is usual to dewater the rougher concentrate and return the water to the primary circuit for two reasons : First, to supply the regrinding mill with a thick enough pulp for efficient operation, and, secondly, as far as possible to prevent the reagents used in the roughing circuit from entering the cleaning section.
In normal practice a recovery of over 90% of the copper which is present as a sulphide is generally possible, whatever the flotation process or circuit employed. As regards the average grade of concentrate, no more can be said than that it depends on the class of the copper-bearing minerals present and their mode of occurrence and on the character of the gangue. It usually contains over 20% of copper, but a difficult chalcopyritic ore may yield a concentrate with less than that percentage, while it is theoretically possible to obtain one running over 75% should the mineral consist entirely of pure chalcocite.
The flotation of native copper ores is nearly always preceded by gravity concentration in jigs and tables not only because the combined process is more economical as regards costs, but also because the copper often occurs as large grains which flatten out during grinding and cannot be broken to a size small enough for flotation. The flow sheet depends on the mode of occurrence of the mineral. The tailings from some of the gravity concentration machines may be low enough in value to be discarded, but those products which still contain too much copper to be sent to waste are thickened and reground, should either operation be necessary, and then floated with pine oil and a xanthate or aerofloat reagent in a neutral or slightly alkaline circuit. The reagent consumption is approximately the same as that given for the treatment of copper- bearing sulphides. While a pine oil, lime, and ethyl xanthate mixture has proved satisfactory, better results have sometimes been obtained by the substitution of aerofloat and sodium di-ethyl-di-thio-phosphate, soda ash being used instead of lime on account of its gangue deflocculating properties. On the average 0-12 lb. per ton of aerofloat and 0.03 lb. of the di-thio-phosphate are substituted for 0.1 lb. of xanthate.
Since a high-grade concentrate is desired in order to keep smelting costs as low as possible, the circuit usually comprises two stages of cleaning. In most plants flotation is carried out in mechanically agitated machines.
The problem of the flotation of oxidized copper ores has not yet been solved. One or two special processes are in operation for the flotation of malachite and azurite, but none of them has more than a limited application; nor has any method been worked out on a large scale for the bulk flotation of mixed oxidized and sulphide copper minerals when the former are present in the ore in appreciable quantity.
Max Feeding size <25mm Discharge size0.075-0.4mm Typesoverflow ball mills, grate discharge ball mills Service 24hrs quotation, custom made parts, processing flow design & optimization, one year warranty, on-site installation.
Ball mill, also known as ball grinding machine, a well-known ore grinding machine, widely used in the mining, construction, aggregate application. JXSC start the ball mill business since 1985, supply globally service includes design, manufacturing, installation, and free operation training. Type according to the discharge type, overflow ball mill, grate discharge ball mill; according to the grinding conditions, wet milling, dry grinding; according to the ball mill media. Wet grinding gold, chrome, tin, coltan, tantalite, silica sand, lead, pebble, and the like mining application. Dry grinding cement, building stone, power, etc. Grinding media ball steel ball, manganese, chrome, ceramic ball, etc. Common steel ball sizes 40mm, 60mm, 80mm, 100mm, 120mm. Ball mill liner Natural rubber plate, manganese steel plate, 50-130mm custom thickness. Features 1. Effective grinding technology for diverse applications 2. Long life and minimum maintenance 3. Automatization 4. Working Continuously 5. Quality guarantee, safe operation, energy-saving. The ball grinding mill machine usually coordinates with other rock crusher machines, like jaw crusher, cone crusher, to reduce the ore particle into fine and superfine size. Ball mills grinding tasks can be done under dry or wet conditions. Get to know more details of rock crushers, ore grinders, contact us!
Ball mill parts feed, discharge, barrel, gear, motor, reducer, bearing, bearing seat, frame, liner plate, steel ball, etc. Contact our overseas office for buying ball mill components, wear parts, and your mine site visits. Ball mill working principle High energy ball milling is a type of powder grinding mill used to grind ores and other materials to 25 mesh or extremely fine powders, mainly used in the mineral processing industry, both in open or closed circuits. Ball milling is a grinding method that reduces the product into a controlled final grind and a uniform size, usually, the manganese, iron, steel balls or ceramic are used in the collision container. The ball milling process prepared by rod mill, sag mill (autogenous / semi autogenous grinding mill), jaw crusher, cone crusher, and other single or multistage crushing and screening. Ball mill manufacturer With more than 35 years of experience in grinding balls mill technology, JXSC design and produce heavy-duty scientific ball mill with long life minimum maintenance among industrial use, laboratory use. Besides, portable ball mills are designed for the mobile mineral processing plant. How much the ball mill, and how much invest a crushing plant? contact us today! Find more ball mill diagram at ball mill PDF ServiceBall mill design, Testing of the material, grinding circuit design, on site installation. The ball grinding mill machine usually coordinates with other rock crusher machines, like jaw crusher, cone crusher, get to know more details of rock crushers, ore grinders, contact us! sag mill vs ball mill, rod mill vs ball mill
How many types of ball mill 1. Based on the axial orientation a. Horizontal ball mill. It is the most common type supplied from ball mill manufacturers in China. Although the capacity, specification, and structure may vary from every supplier, they are basically shaped like a cylinder with a drum inside its chamber. As the name implies, it comes in a longer and thinner shape form that vertical ball mills. Most horizontal ball mills have timers that shut down automatically when the material is fully processed. b. Vertical ball mills are not very commonly used in industries owing to its capacity limitation and specific structure. Vertical roller mill comes in the form of an erect cylinder rather than a horizontal type like a detachable drum, that is the vertical grinding mill only produced base on custom requirements by vertical ball mill manufacturers. 2. Base on the loading capacity Ball mill manufacturers in China design different ball mill sizes to meet the customers from various sectors of the public administration, such as colleges and universities, metallurgical institutes, and mines. a. Industrial ball mills. They are applied in the manufacturing factories, where they need them to grind a huge amount of material into specific particles, and alway interlink with other equipment like feeder, vibrating screen. Such as ball mill for mining, ceramic industry, cement grinding. b. Planetary Ball Mills, small ball mill. They are intended for usage in the testing laboratory, usually come in the form of vertical structure, has a small chamber and small loading capacity. Ball mill for sale In all the ore mining beneficiation and concentrating processes, including gravity separation, chemical, froth flotation, the working principle is to prepare fine size ores by crushing and grinding often with rock crushers, rod mill, and ball mils for the subsequent treatment. Over a period of many years development, the fine grinding fineness have been reduced many times, and the ball mill machine has become the widest used grinding machine in various applications due to solid structure, and low operation cost. The ball miller machine is a tumbling mill that uses steel milling balls as the grinding media, applied in either primary grinding or secondary grinding applications. The feed can be dry or wet, as for dry materials process, the shell dustproof to minimize the dust pollution. Gear drive mill barrel tumbles iron or steel balls with the ore at a speed. Usually, the balls filling rate about 40%, the mill balls size are initially 3080 cm diameter but gradually wore away as the ore was ground. In general, ball mill grinder can be fed either wet or dry, the ball mill machine is classed by electric power rather than diameter and capacity. JXSC ball mill manufacturer has industrial ball mill and small ball mill for sale, power range 18.5-800KW. During the production process, the ball grinding machine may be called cement mill, limestone ball mill, sand mill, coal mill, pebble mill, rotary ball mill, wet grinding mill, etc. JXSC ball mills are designed for high capacity long service, good quality match Metso ball mill. Grinding media Grinding balls for mining usually adopt wet grinding ball mills, mostly manganese, steel, lead balls. Ceramic balls for ball mill often seen in the laboratory. Types of ball mill: wet grinding ball mill, dry grinding ball mill, horizontal ball mill, vibration mill, large ball mill, coal mill, stone mill grinder, tumbling ball mill, etc. The ball mill barrel is filled with powder and milling media, the powder can reduce the balls falling impact, but if the power too much that may cause balls to stick to the container side. Along with the rotational force, the crushing action mill the power, so, it is essential to ensure that there is enough space for media to tumble effectively. How does ball mill work The material fed into the drum through the hopper, motor drive cylinder rotates, causing grinding balls rises and falls follow the drum rotation direction, the grinding media be lifted to a certain height and then fall back into the cylinder and onto the material to be ground. The rotation speed is a key point related to the ball mill efficiency, rotation speed too great or too small, neither bring good grinding result. Based on experience, the rotat
ion is usually set between 4-20/minute, if the speed too great, may create centrifuge force thus the grinding balls stay with the mill perimeter and dont fall. In summary, it depends on the mill diameter, the larger the diameter, the slower the rotation (the suitable rotation speed adjusted before delivery). What is critical speed of ball mill? The critical speed of the ball mill is the speed at which the centrifugal force is equal to the gravity on the inner surface of the mill so that no ball falls from its position onto the mill shell. Ball mill machines usually operates at 65-75% of critical speed. What is the ball mill price? There are many factors affects the ball mill cost, for quicker quotations, kindly let me know the following basic information. (1) Application, what is the grinding material? (2) required capacity, feeding and discharge size (3) dry or wet grinding (4) single machine or complete processing plant, etc.
Grinding circuits are fed at a controlled rate from the stockpile or bins holding the crusher plant product. There may be a number of grinding circuits in parallel, each circuit taking a definite fraction of the feed. An example is the Highland Valley Cu/Mo plant with five parallel grinding lines (Chapter 12). Parallel mill circuits increase circuit flexibility, since individual units can be shut down or the feed rate can be changed, with a manageable effect on production. Fewer mills are, however, easier to control and capital and installation costs are lower, so the number of mills must be decided at the design stage.
The high unit capacity SAG mill/ball mill circuit is dominant today and has contributed toward substantial savings in capital and operating costs, which has in turn made many low-grade, high-tonnage operations such as copper and gold ores feasible. Future circuits may see increasing use of high pressure grinding rolls (Rosas et al., 2012).
Autogenous grinding or semi-autogenous grinding mills can be operated in open or closed circuit. However, even in open circuit, a coarse classifier such as a trommel attached to the mill, or a vibrating screen can be used. The oversize material is recycled either externally or internally. In internal recycling, the coarse material is conveyed by a reverse spiral or water jet back down the center of the trommel into the mill. External recycling can be continuous, achieved by conveyor belt, or is batch where the material is stockpiled and periodically fed back into the mill by front-end loader.
In Figure 7.35 shows the SAG mill closed with a crusher (recycle or pebble crusher). In SAG mill operation, the grinding rate passes through a minimum at a critical size (Chapter 5), which represents material too large to be broken by the steel grinding media, but has a low self-breakage rate. If the critical size material, typically 2550mm, is accumulated the mill energy efficiency will deteriorate, and the mill feed rate decreases. As a solution, additional large holes, or pebble ports (e.g., 40100mm), are cut into the mill grate, allowing coarse material to exit the mill. The crusher in closed circuit is then used to reduce the size of the critical size material and return it to the mill. As the pebble ports also allow steel balls to exit, a steel removal system (such as a guard magnet, Chapters 2 and 13Chapter 2Chapter 13) must be installed to prevent them from entering the crusher. (Because of this requirement, closing a SAG mill with a crusher is not used in magnetic iron ore grinding circuits.) This circuit configuration is common as it usually produces a significant increase in throughput and energy efficiency due to the removal of the critical size material.
An example SABC-A circuit is the Cadia Hill Gold Mine, New South Wales, Australia (Dunne et al., 2001). The project economics study indicated a single grinding line. The circuit comprises a SAG mill, 12m diameter by 6.1m length (belly inside liners, the effective grinding volume), two pebble crushers, and two ball mills in parallel closed with cyclones. The SAG mill is fitted with a 20MW gearless drive motor with bi-directional rotational capacity. (Reversing direction evens out wear on liners with symmetrical profile and prolongs operating time.) The SAG mill was designed to treat 2,065t h1 of ore at a ball charge of 8% volume, total filling of 25% volume, and an operating mill speed of 74% of critical. The mill is fitted with 80mm grates with total grate open area of 7.66m2 (Hart et al., 2001). A 4.5m diameter by 5.2m long trommel screens the discharge product at a cut size of ca. 12mm. Material less than 12mm falls into a cyclone feed sump, where it is combined with discharge from the ball mills. Oversize pebbles from the trommel are conveyed to a surge bin of 735t capacity, adjacent to the pebble crushers. Two cone crushers with a closed side set of 1216mm are used to crush the pebbles with a designed product P80 of 12mm and an expected total recycle pebble rate of 725t h1. The crushed pebbles fall directly onto the SAG mill feed belt and return to the SAG mill.
SAG mill product feeds two parallel ball mills of 6.6m11.1m (internal diameterlength), each with a 9.7MW twin pinion drive. The ball mills are operated at a ball charge volume of 3032% and 78.5% critical speed. The SAG mill trommel undersize is combined with the ball mills discharge and pumped to two parallel packs (clusters) of twelve 660mm diameter cyclones. The cyclone underflow from each line reports to a ball mill, while the cyclone overflow is directed to the flotation circuit. The designed ball milling circuit product is 80% passing 150m.
Several large tonnage copper porphyry plants in Chile use an open-circuit SAG configuration where the pebble crusher product is directed to the ball mills (SABC-B circuit). The original grinding circuit at Los Bronces is an example: the pebbles generated in the two SAG mills are crushed in a satellite pebble crushing plant, and then are conveyed to the three ball mills (Mogla and Grunwald, 2008).
Hydrocyclones have come to dominate classification when dealing with fine particle sizes in closed grinding circuits (<200m). However, recent developments in screen technology (Chapter 8) have renewed interest in using screens in grinding circuits. Screens separate on the basis of size and are not directly influenced by the density spread in the feed minerals. This can be an advantage. Screens also do not have a bypass fraction, and as Example 9.2 has shown, bypass can be quite large (over 30% in that case). Figure 9.8 shows an example of the difference in partition curve for cyclones and screens. The data is from the El Brocal concentrator in Peru with evaluations before and after the hydrocyclones were replaced with a Derrick Stack Sizer (see Chapter 8) in the grinding circuit (Dndar et al., 2014). Consistent with expectation, compared to the cyclone the screen had a sharper separation (slope of curve is higher) and little bypass. An increase in grinding circuit capacity was reported due to higher breakage rates after implementing the screen. This was attributed to the elimination of the bypass, reducing the amount of fine material sent back to the grinding mills which tends to cushion particleparticle impacts.
Circulation of material occurs in several parts of a mineral processing flowsheet, in grinding and flotation circuits, for example, as well as the crushing stage. In the present context, the circulating load (C) is the mass of coarse material returned from the screen to the crusher relative to the circuit final product (or fresh feed to the circuit), often quoted as a percentage. Figure 8.2 shows two closed circuit arrangements. Circuit (a) was considered in Chapter 6 (Example 6.1), and circuit (b) is an alternative. The symbols have the same meaning as before. The relationship of circulating load to screen efficiency for circuit (a) was derived in Example 6.1, namely (where all factors are as fractions):
The circulating load as a function of screen efficiency for the two circuits is shown in Figure 8.3. The circulating load increases with decreasing screen efficiency and as crusher product coarsens (f or r decreases), which is related to the crusher set (specifically the closed side setting, c.s.s.). For circuit (a) C also increases as the fresh feed coarsens (n decreases), which is likely coming from another crusher. In this manner, the circulating load can be related to crusher settings.
In industrial grinding process, in addition to goal of productivity maximization, other purposes of deterministic grinding circuit optimization have to satisfy the upper bound constraints on the control variables. We know that there lies a tradeoff between the throughput (TP) and the percent passing of midsize classes (MS) from the previous work of Mitra and Gopinath,2004. In deterministic optimization formulation, there are certain parameters which we will assume them as constant. But, in real life that may not be case. There are such six parameters in our industrial grinding process which are R, B, R, B are the grindability indices and grindability exponents for the rod mill (RMGI) and the ball mill (BMGI); and P, S are the sharpness indices for the primary (PCSI) and secondary cyclones (SCSI). These parameters are treated as constant in deterministic formulation. As they are going to be treated as uncertain parameters in the OUU formulation. These parameters are assumed uncertain because most of them are obtained from the regression of experimental data and thus are subject to uncertainty due to experimental and regression errors. In the next part of the section, we consider them as fuzzy numbers and solve the OUU problem by FEVM. In FEVM formulation, the uncertain parameters are considered as fuzzy numbers and the uncertain formulation is transformed into the deterministic formulation by expectation calculations for both objective function and constraints. So, the converted deterministic multi-objective optimization problem is expressed as:
Another spinning batch concentrator (Figure 10.27), it is designed principally for the recovery of free gold in grinding circuit classifier underflows where, again, a very small (<1%) mass pull to concentrate is required. The feed first flows up the sides of a cone-shaped bowl, where it stratifies according to particle density before passing over a concentrate bed fluidized from behind by back-pressure (process) water. The bed retains dense particles such as gold, and lighter gangue particles are washed over the top. Periodically the feed is stopped, the bed rinsed to remove any remaining lights and is then flushed out as the heavy product. Rinsing/flushing frequency, which is under automatic control, is determined from grade and recovery requirements.
The units come in several designs, the Semi-Batch (SB), Ultrafine (UF), and i-Con, designed for small scale and artisanal miners. The first installation was at the Blackdome Gold Mine, British Columbia, Canada, in 1986 (Nesset, 2011).
These two batch centrifugal concentrators have been widely applied in the recovery of gold, platinum, silver, mercury, and native copper; continuous versions are also operational, the Knelson Continuous Variable Discharge (CVD) and the Falcon Continuous (C) (Klein et al., 2010; Nesset, 2011).
To liberate minerals from sparsely distributed and depleting the ore bodies finer grinding than generally obtained by the conventional Rod Mill Ball Mill grinding circuits is needed. Longer grinding periods in the conventional milling processes prove too expensive mainly due to large power consumption. Stirrer mills have been tried in mineral industry with considerable success and have therefore been increasingly used. In this chapter, the theories involved in the design and operation of these mills, as established till now, are explained. Further theoretical studies and designs of the mills are still in progress for a better understanding and improved operation. Presently, the mills have been proved to be economically viable and the mineral of interest conducive to improved recovery and grade.
IMP Technologies Pty. Ltd. has recently tested a pilot-scale super fine crusher that operates on dry ore and is envisaged as a possible alternative to fine or ultra-fine grinding circuits (Kelsey and Kelly, 2014). The unit includes a rotating compression chamber and an internal gyrating mandrel (Figure 6.13). Material is fed into the compression chamber and builds until the gyratory motion of the mandrel is engaged. Axial displacement of the compression chamber and the gyratory motion of the mandrel result in fine grinding of the feed material. In one example, a feed F80 of 300m was reduced to P80 of 8m, estimated to be the equivalent to two stages of grinding. This development is the latest in a resurgence in crushing technology resulting from the competition of AG/SAG milling and the demands for increased comminution energy efficiency.
The iron oxide crystal grains in most iron ores are not evenly distributed in size. Spiral separators can therefore be used to take out the coarser iron oxide grains in the primary grinding circuit to save grinding energy and help achieve a higher iron recovery. Figure 9.14 presents a typical flow sheet for processing an oxidized ore containing about 30% Fe using a combination of spiral and SLon magnetite separators and reverse flotation. This ore is mainly composed of hematite, magnetite, and quartz, and the iron oxide crystals range in size from 0.005 to 1.0mm with an average size of about 0.05mm. The average size of the quartz crystals is approximately 0.085mm.
In the primary grinding stage of the flow sheet in Figure 9.14, the ore is first ground down to about 60% -75m and then classified into two size fractions, a coarse size fraction and a fine size fraction. The coarse size fraction is treated with spiral separators to recover part of the final iron ore concentrate. Then, drum LIMS and SLon magnetic separators are used to reject some of the coarse gangue minerals as final tailings. The magnetic products from the LIMS and SLon are sent back to the secondary ball mill for regrinding, and the milled product returns to the primary cyclone classifier.
The fine size fraction is about 90% -75m and is processed using drum LIMS separators and SLon magnetic separators in series to take out the magnetite and hematite, respectively. The magnetic products from the magnetic separators are mixed to generate the feed for reverse flotation to produce another component of the final iron ore concentrate.
The key advantage of this flow sheet lies in the fact that the spirals and SLon magnetic separators take out about 20% of the mass of the final iron concentrate and about 20% of the mass of the final tailings, respectively, from the coarse size fraction. This greatly reduces the masses being fed to the secondary ball mill and reverse flotation, thereby greatly reducing the total processing cost. From the plant results for this flow sheet, an iron concentrate containing 67.5% Fe could be produced from a run-of-mine ore containing 30.1% Fe, at a mass yield to the iron concentrate of 34.9%, an iron recovery of 78.0%, and a tailings grade of 10.2% Fe.
The first step of physical beneficiation is crushing and grinding the iron ore to its liberation size, the maximum size where individual particles of gangue are separated from the iron minerals. A flow sheet of a typical iron ore crushing and grinding circuit is shown in Figure 1.2.2 (based on Ref. ). This type of flow sheet is usually followed when the crude ore contains below 30% iron. The number of steps involved in crushing and grinding depends on various factors such as the hardness of the ore and the level of impurities present .
Jaw and gyratory crushers are used for initial size reduction to convert big rocks into small stones. This is generally followed by a cone crusher. A combination of rod mill and ball mills are then used if the ore must be ground below 325 mesh (45m). Instead of grinding the ore dry, slurry is used as feed for rod or ball mills, to avoid dusting. Oversize and undersize materials are separated using a screen; oversize material goes back for further grinding.
Typically, silica is the main gangue mineral that needs to be separated. Iron ore with high-silica content (more than about 2%) is not considered an acceptable feed for most DR processes. This is due to limitations not in the DR process itself, but the usual customer, an EAF steelmaking shop. EAFs are not designed to handle the large amounts of slag that result from using low-grade iron ores, which makes the BF a better choice in this situation. Besides silica, phosphorus, sulfur, and manganese are other impurities that are not desirable in the product and are removed from the crude ore, if economically and technically feasible.
While used sometimes on final concentrates, such as Fe concentrates, to determine the Blaine number (average particle size deduced from surface area), and on tailings for control of paste thickeners, for example, the prime application is on cyclone overflow for grinding circuit control (Kongas and Saloheimo, 2009). Control of the grinding circuit to produce the target particle size distribution for flotation (or other mineral separation process) at target throughput maximizes efficient use of the installed power.
Continuous measurement of particle size in slurries has been available since 1971, the PSM (particle size monitor) system produced then by Armco Autometrics (subsequently by Svedala and now by Thermo Gamma-Metrics) having been installed in a number of mineral processing plants (Hathaway and Guthnals, 1976).
The PSM system uses ultrasound to determine particle size. This system consists of three sections: the air eliminator, the sensor section, and the electronics section. The air eliminator draws a sample from the process stream and removes entrained air bubbles (which otherwise act as particles in the measurement). The de-aerated pulp then passes between the sensors. Measurement depends on the varying absorption of ultrasonic waves in suspensions of different particle sizes. Since solids concentration also affects the absorption, two pairs of transmitters and receivers, operating at different frequencies, are employed to measure particle size and solids concentration of the pulp, the processing of this information being performed by the electronics. The Thermo GammaMetrics PSM-400MPX (Figure 4.18) handles slurries up to 60% w/w solids and outputs five size fractions simultaneously.
Other measurement principles are now in commercial form for slurries. Direct mechanical measurement of particle size between a moving and fixed ceramic tip, and laser diffraction systems are described by Kongas and Saloheimo (2009). Two recent additions are the CYCLONEtrac systems from CiDRA Minerals Processing (Maron et al., 2014), and the OPUS ultrasonic extinction system from Sympatec (Smith et al., 2010).
CiDRAs CYCLONEtrac PST (particle size tracking) system comprises a hardened probe that penetrates into the cyclone overflow pipe to contact the stream and effectively listens to the impacts of individual particles. The output is % above (or below) a given size and has been shown to compare well with sieve sizing (Maron et al., 2014). The OPUS ultrasonic extinction system (USE) transmits ultrasonic waves through a slurry that interact with the suspended particles. The detected signal is converted into a particle size distribution, the number of frequencies used giving the number of size classes measured. Applications on ores can cover a size range from 1 to 1,000m (Smith et al., 2010).
In addition to particles size, recent developments have included sensors to detect malfunctioning cyclones. Westendorf et al. (2015) describe the use of sensors (from Portage Technologies) on cyclone overflow and underflow piping. CiDRAs CYCLONEtrac OSM (oversize monitor) is attached to the outside of the cyclone overflow pipe and detects the acoustic signal as oversize particles (rocks) hit the pipe (Cirulis and Russell, 2011). The systems are readily installed on individual cyclones thus permitting poorly operating units to be identified and changed while allowing the cyclone battery to remain in operation. Figure 4.19 shows an installation of both CiDRA systems (PST, OSM) on the overflow pipe from a cyclone.
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Several years ago, Davis assumed that the rate of wear of the different sizes of balls in a ball mill was directly proportional to the weight of each ball, and he evolved a formula for calculating a balanced charge. Operators have used this formula when purchasing balls for a new mill or when reloading an old one that had been emptied for repair. The formula required that the largest ball size and the size to be rejected should be determined, and after that the other sizes were set. Stress was laid on the coarsest size, and to facilitate the use of the formula many writers have made their contribution by reporting ratio of coarsest particle size to the optimum ball size. Close adherence to this ratio has prevented giving attention to sizes and amounts of particles not falling in the category of the coarsest size.
The inadequacy of the formula and the futility of extensive experimentation for ratio determinations involving the coarsest particle size only is at once obvious when it is seen that the formula did not take into account the slow grinding rate of the finer sizes of ore and the amount present. To be sure, operators who were doing very fine grinding have sometimes altered the make-up load by using some additional small balls with the big ones, but this practice has been somewhat haphazard. Too much of the work has followed the old idea that there should be no ball present that is incapable of crushing the largest particle in the feed.
Today operators have a keener sense of the relatively large amount of work required to finish the finest sizes, so that the insufficiency of the formula is readily seen. It would have been fortunate had the formula been devised to attract more attention to the large amount of finer but unfinished particles. The formula is excellent from the basis of balance with respect to ball wear, but the literature has contained very little about the rationing of ball sizes for the best grinding of all sizes and amounts of particles extending throughout the length of the mill. Research has submitted in this matter.
It is not denied that the coarse particles have to be crushed else no fine material would accrue, but here the fact is emphasized that when crushing to 200-mesh stress should be on the selection of balls of the right size and amount to crush, say, from 100- to 200-mesh; or, when crushing to 65-mesh, the operator should judiciously load the mill for crushing from 48- to 65-mesh. If this were done, the circulating load would be relieved of the large amount of nearly finished size, and in its stead there would be some coarser material from which the classifier could more easily remove the finished size. Opposing this idea is the fact that a coarse circulating load would be undesirable in some of the recent supplementary recovery processes. However, this objection might be met by introducing a bypassing screen at the end of the ball mill.
Tests of other experimenters have been supplemented with detailed information on the optimum size of balls for grinding sized ore. Figures have been obtained that show what particular size of ball is the most efficient in crushing certain sizes of chert and dolomite. It is fortunate that this work has been done,
because it has brought out facts that would have been unsuspected otherwise. The method used here for showing what particular size of ball is best for a particular particle size of ore is to some degree unique. The reason for this is that usually such tests have been run to finish the grinding at a fine size. Those tests were as much a criterion of the work on the particle size in the finished product as of the feed, but they were not so interpreted. The tests reported in this paper are different because the first step in reduction is given the main emphasis.
As a guide in laying out this work, a mill was visualized as divided into sections. The first section had the largest media and performed the first step in grinding by reducing the particles for the second section; the second section, in turn, used smaller media to reduce the articles for the third section; and so on. This line of thought was the basis for the distribution of sizes in the ball loads already mentioned.
The ball sizes were 2.75-to 0.62-inch and the ore sizes plus 65- to plus 10-mesh. The results for chert are shown in four series in table 27 and for dolomite in five series in table 28. The ore (feed) sizes are in quotation marks because they are only nominal; their meaning is set forth in the sizing analyses under feed.
Any plan adopted would give but litle more than an approximation of the facts sought, owing to the difficulty in timing the grinding correctly. If it is desired to find the effect of balls grinding 20-mesh ore and the mill is loaded with 20-mesh material, the grinding time should be infinitely short, because fine particles are made as soon as the mill starts and if the run continues the test is of the comminuted products of the 20-mesh sample rather than that which was supplied for the test.
Extrapolation back to zero time would be desirable if it were possible. However, a very short period is unsatisfactory because the flaky particles, being the first to yield, would give a wrong impression of the sample as a whole. Long grinding periods would be useless because the particle size at the end of the run would be too far removed from the original particle size under investigation. A mean procedure had to be adopted.
The surface calculations that are given must be used guardedly, else they will be misleading. The fine particle sizes are likely to be weighted too much; when the ball size for crushing 10-mesh sizes through 14-mesh is sought, the very fine sizes should be weighted with caution.
A casual examination of each series for minimum of cumulative weights in the coarse sizes of the screen analyses probably would be a fair guide to the best ball size. But this minimum, though important of consideration, is not final, because the amount, power, and time have to be taken into account. These three quantities are resolved into tons per horsepower-hour and will be applied in table 29. Before going to that table, however, the present tables may be used to bring out a fact not commonly knownballs that were too large as well as balls that were too small failed in selective grinding. In any of the series except the last one of each table, where the largest ball sizes were not large enough, the low cumulative percentage weight of the coarse sizes is in a mean position and rises with the use of larger as well as smaller balls. Hence, it is shown that balls that were too large did nonselective grinding.
mesh size in table 27 and are shown in sizing diagrams. The percentage weights of the products from the largest, and the smallest balls are shown by broken lines. They are high in the upper part of the diagram. Their position shows that much of the coarse material was not reduced through 35-mesh. The solid line shows good selective work of the balls of optimum size.
In the study of these diagrams, it must be remembered that the main variables in the tests were ball size and that the tests were timed to give the same amount of subsieve size. The conditions imposed on the tests were entirely different from closed-circuit grinding, in which the composite feeds would have been unlike, although the new feeds might have been the same.
The nonselective grinding of the off-size balls may be explained as follows: The largest balls failed on the coarsest sizes because they did not offer a sufficient number of points of contact for the number of grains present; hence, some of the particles remained at the end of the test. Furthermore, due to the small number of points of contact of such large balls, the crushing impulse was so great that the grains that did meet it received excessive comminution and much of the subsieve size resulted.
The smallest balls had so many points of contact that the impulse at a given point was too much reduced to exert sufficient stress on the coarsest particles; hence, some of them remained without the desired reduction. However, a few that were reduced yielded grains readily comminuted by the smallest balls, and much subsieve size again resulted; hence, there was an intermediate ball size for the best work.
Table 29 will now be discussed: It is made by using the two preceding tables. It gives the amount of the coarsest size per unit of power crushed through a stated coarse but finer size. To illustrate the method of calculation, take the first test in table 27: The amount of plus 65-mesh crushed through 100-mesh is 89.563.3=26.2 parts per hundred, and by the table the ton per horsepower-hour was 0.16; hence, the tons per horsepower-hour crushed through 100-mesh was 26.2/1000.16=0.042. Similarly, in the first test in the second series 97.352.0=45.3, and 45.3/1000.186=0.084 ton per horsepower-hour through 48-mesh. Thus, table 29 has four series of tests or chert and five series for dolomite. The preferred value in each series is underscored to show what seems to be the preferable ball size. The optimum ball size for grinding closely sized particles through the limiting screen, as determined by these experiments, may be expressed in the following equation
where D is diameter of ball, d is diameter of particle to be ground, and K is a constant depending on the grindability of the ore. When D and d are expressed in inches, the value of K for chert is 55 and for dolomite is 35. This formula is of the same type as that developed by Starke. He evaluated the grind through a broader range and his dimensions are in microns.
Having selected the best ball size, it will be seen by referring to tables 27 and 28 that the preferable ball size usually gave the best capacity and efficiency. Also, the preferable ball size coincides closely with the best selective grinding, the main exception being the plus 10-mesh series in table 28. There the preferable ball size is smaller than the size for the best selective grinding. Probably the exception is due to an error in planning the plus 10-mesh series; the time periods were too long and too much grinding resulted. The spread in reduction in this series was greater than in any other series. It was intended to avoid such a broad spread in reduction. In the study of the exception and the study of the sizing analyses in the other tests an attempt has been made to gain additional information by using the Gaudin log-log method for plotting sizinganalyses, but the results were not satisfactory. It is believed, however, that the method was not expected to apply to the moderate reduction of a sized product.
Tables 27 and 28 cannot be dismissed without consideration of the variation of power throughout a test. Figure 5 is submitted for that purpose. In it the time extends from 0 to 3.5 minutes. The change in power through the grinding periods was watched in all the tests. This change is illustrated in figure 5, which deals with the plus 20-mesh size in table 28. In the discussion of this figure, what will be said about the relation of power to other factors is premised by the belief that the degree to which the balls nip the particles influences the power, and that when nipping is best the power will be the highest. The curve at the bottom of the figure shows that the 2.75-inch balls required less power than the other loads. The balls were too big for good nipping, and as the grinding continued they became relatively bigger and further power reduction resulted. Correlated with this is the fact that the grinding was poor in selection and unsatisfactory in capacity and efficiency. (It is not consistent to compare the numerical-values of capacities and efficiencies of one series in tables 27 and 28 with those of another series. The principles underlying the reason were mentioned under Sillimanite balls.)
Turning next to the deportment of the 0.62-inch balls, which were the smallest in the group, the change in power from beginning to end of the run is in a reverse order from that with the largest balls. The balls were too small for good nipping, but as comminution proceeded they became relatively larger so that nipping and power increased but did not reach the high power indicating good nipping. The selective grinding, capacity, and efficiency were again poor.
The record of the 1-inch balls is more favorable. The power was high throughout the test, indicating that a desirable mean size had been reached. The selective grinding, capacities, and efficiencies were good. This all indicates that when nipping is best the mill (when not run too fast) will do its best work. This statement is not new; the evidence is given for those who wish to weigh it.
A comprehensive examination of mills that segregate the ball sizes shows that they require mixtures containing a greater number of small balls than is supplied by the Davis ball load. This deficiency was met by using the rationed ball load, in which small balls predominated. Before going ahead, the mills will be considered.
Conical mills and cylindrical mills with grids were contemplated in introducing the new loads. Hence, these mills must be discussed before showing the tests, and they must be compared with the standard cylindrical mill.
Should the ball sizes be segregated, or should they be mixed as in the standard cylindrical mill? In the metallics industry the most effective method of segregating is to place the mills in series and use succeedingly smaller balls from first to last mill in the series. In the cement industry, dividers or grids are used to divide the long mills into sections, each of which has the appropriate size of medium. Finished material is removed at the end of each section.
With the knowledge that the cone of a conical mill functions like a grid in segregating the balls with respect to size, conical mills were built and tested. The first one was only 3 feet long. A taper of 2 inches to the foot was ample to segregate the largest balls in the big end and the smallest balls in the small end. Grinding tests in this mill with a rationed ball load were compared with the old cylindrical mill loaded with the old style ball load. A decided advantage was gained by the newer practice.
A larger conical mill was built and is shown in figure 6. It was 6 feet long and had the same taper as the smaller one. The big end was 2 feet in diameter and the small end 1 foot. The ability of the mill to segregate the balls was demonstrated by tests.
Grinding tests with several types of mills and ball loads led to the conclusion that advantages that had been gained were due more to the appropriate average size of balls than to the new design of mill. It was difficult to show that the conical mills had an outstanding advantage over the cylindrical mill. The 6-foot conical mill had a disadvantage; it induced the media to drift to the big end and pile up there so much that the balls passed through the feed entrance into the scoop. For a simple remedy a grid was placed on the feed opening to retain the load. A change was made to a cylindrical mill lined with a series of identical truncated cones. The idea was suggested by C. L. Carman, of Independence, Kans.
Although the efficiency of the long cone was good, the loss in capacity induced by the taper was marked. This may be shown by the following analysis: If the last unit section with diameter D2 = 1 foot could be speeded up to the same percent critical as the first unit section with diameter D1 = 2 feet, it would have a relatively low capacity
A 2- by 3-foot cylindrical mill was lined to employ the conical effect, but instead of having one cone it had three identical truncated cones, end to end, and apexing in the same direction. (See fig. 7.) Any cylindrical mill may be lined in this fashion by using liners tapered in thickness. If the liners are 2 feet long, a 12-foot mill would have six truncated cones, end to end. The mill would have the same capacity at the discharge end as at the feed end. In some way, at least, this would be an advantage over the long cone. The mill with truncated cones proved to be as good a sizing device as the long single cone, but when compared with the old cylindrical mill its advantage as a grinder was not marked.
Finally, a cylindrical mill with a grid was used. The grid was 1 foot from the feed end of a 2- by 3-foot cylindrical mill. Balls of 2.5 inches to 1 inch were placed in the feed-end sections and 0.75-inch balls in the discharge end. The grinding was moderately better than without the grid. Possibly the grid would have appeared to have more advantage if the feed had been coarser and the finishing finer. The
In table 30, grinding to a fine size was stressed to give the extra amount of small media in the new load a chance to work advantageously. Grinding was continuous and about 74 percent of the product passed through 200-mesh. The Davis ball load in the cylindrical mill was used first; next, the rationed ball load was used in the same mill; and finally, the rationed ball load was used in a mill having a lining of truncated cones. In selecting the Davis ball load the no. 1 load was used instead of no. 2 in accordance with the old idea that all of the balls should be of a size to crush any of the particles of ore. The free migration of the ore induced by the large, interstices would be compatible with a heavy circulating load. In the cylindrical mill the work of the rationed ball load was about 60 percent better than the Davis ball load, and when the mill which was lined with truncated cones was used there was a further gain of about 5 percent. The rationed ball load left more of the coarse sizes unfinished.
In table 31 the results of five tests with different ball loads in cylindrical, grid, and conical mills are shown. The feed was coarser than was used in table 30. The grinding in test 2 with the rationed ball load, which contained 64 percent of 0.75-inch balls, was about 44 percent more efficient than with the Davis load. The power was about 11 percent higher. If Davis ball load no. 2 instead of no. 1 had been used, the divergence in grinding results would have been reduced. In test 3, in which the grid was used to segregate the different sizes of balls, a further advantage of about 4 percent in efficiency is shown. The conical mill in test 4 increased the efficiency to 58 percent more than in test 1. The efficiency with the long (6-foot) conical mill was about the same as with the short (3-foot) one.
The validity of having graded sizes of balls to grind the ore in steps with ample provisions for a circulating load and removal of fines in each step cannot be denied, but without this quick removal of finished material the advantage was not great.
In the conical mills or in the grid mill, as used in these tests, it was difficult to set a correct feed rate. If the feed were too fast some of the coarse particles would pass the zone intended to grind them. Having passed that zone, they were likely to continue without being ground. Again, if the feed were too slow, energy would be wasted by making the fine particles remain too long with the coarse medium. Nonselective grinding and inefficiency would result.
Thus far the evidence of the efficacy of a rationed load in plant operation may be questioned because, as is shown by tables 30 and 31, the coarse sizes were not reduced as much as with the Davis load. Fear was entertained lest a circulating load might develop trouble- some characteristics. Hence, closed-circuit grinding was tried.
Rationed ball sizes were of advantage in batch and open-circuit grinding. The degree depended on the particle size of feed and product. Examination will now be made to see if the deportment of rationed sizes is satisfactory in closed-circuit grinding.
The tests were performed as shown in tables 32 and 33. In the first table dolomite B was used, and in the second the feed was chert rejects from earlier grinding tests. The procedures in the two tables have one fundamental difference; in table 32 the feed to the rationed ball load was increased on account of the extra efficiency of the rationed sizes, whereas in table 33 the feed was maintained at the same rate but the mill speed was reduced. That is, in the first table the advantage is shown by the increased amount of ore ground, and in the second the advantage is shown by the power saved. If preference is given to one of the two methods it should apply to the latter, because in it the two ball loads being compared deal with the same amount of feed, and the drag is worked under almost identical conditions. The pulp consistencies of the drag overflows were maintained at 17 percent solids.
In table 32 the drag classifier finished at a finer size when rationed sizes of balls were used. This variation is on the right side for safe conclusions about the advantage of the new ball load. With the Davis ball load, 2.95 pounds per minute were finished, and with the rationed ball load the amount was increased to 4.06 pounds perminute an increase of 37 percent. The surface tons per hour show, an increase of 45 percent in favor of the rationed ball sizes, and the surface tons per horsepower-hour show a more moderate advantage37 percent. The reason the advantage in capacity was greater than in efficiency is because of the difference in power in the two tests; the smaller balls required more power than the larger ones. The surface calculations are made from the part of the table marked section 3. There a composite feed has been calculated, so that surface calculation can be based on feed and product. However, the ultimate values would have been the same if the sizing analyses of new feed and over- flow in sections 1 and 2 had been used.
It will be seen that the circulating loads in each table are about the same, respectively. Due precaution was taken to make sure that the circulating load was balanced, about 2 hours being required after the last adjustment.
The closed-circuit set-ups are shown in figures 8 and 9. They do not include the inclined belt and weightometer formerly used. A better plan was to permit the drag sand to fall into buckets and at set intervals to pass the sand back to the new-ore belt feeder after a hurried weighing. The record of the weights obtained after decanting
superfluous water indicated the trend of the circulating load, but a more accurate estimate was made at frequent intervals by catching the ball-mill discharge in a graduate and weighing it. The weight of solids minus new feed gave the circulating load with exactness. The test was continued for a goodly period after the amount of discharge became constant.
In the two tests shown in Table 33, the overflows are nearly identical. The innovation in the manner of conducting the tests, as stated before, was to keep the new feed constant and reduce the speed of the mill containing the new ball load until the circulating load in section 2 was the same as in section 1. When the new ball load was used, the speed was reduced from 70 to 55 percent critical and the capacity was maintained. The increase in efficiency was 28 percent. The Davis ball load took 22.6 percent more power than the rationed ball load.
The comparison of different sizes of media when the mill speeds are not the same might not have been justified by the old literature, but it is justified by table 13, which shows that for speeds from 40 to 70 percent critical, inclusive, the efficiencies were almost identical when the amount of ore in the mill was the same; of course, capacity increased with speed. It is readily seen from table 33 that the capacity with the rationed ball load at 55 percent speed was about the same as with the Davis load at 70 percent speed. If the finishing could have been at 200-mesh in all the closed-circuit tests, the load of large balls would have been greatly handicapped and the load of small balls would have had a greater relative advantage. Then the difference in efficiency might have been as much as 75 percent. The grinding seems to have been a little more selective with the larger media.
By table 2 the diameter of the ball of average weight in the rationed load no. 2 was 1 inch. A load of 1-inch balls would have given about the same results but would not have permitted the study of the effect of segregation in the grid and conical mills. Furthermore, the practical application would have been doubtful. A Davis ball load with sizes from 1 to 1 inch would have done good work, but it would not have been representative of the old standard because some of the balls would have been too small to crush the largest particles.
The quantities obtained in these tests enable the mill man to get a vision of the amount of power required to do his grinding. Take, for example, the tests represented by section 1 in table 33, in which grinding was to flotation size by what may be called the ordinary ball load and the ore feed was almost 100-percent Tri-State chert through 8-mesh. Calculations show that the net energy input was 21 horsepower-hours per ton. One-third should be added for friction and motor losses, which would bring the motor input up to 28 horsepower-hours per ton of ore. An ore would have to be rich to justify the expenditure of so much additional power for grinding.
Milling circuit and ball size distribution models determined the make-up ball charge.Investigations focused on mono and binary mixture of ball sizes for make-up charge.The results found reccommended charging mono-sized balls in the make-up charge.Optimal conditions for the size and composition of make-up charge were determined.The conditions enable maximum production of the desired size range for flotation.
Ball size distribution is commonly used to optimise and control the quality of the mill product. A simulation model combining milling circuit and ball size distribution was used to determine the best make-up ball charge. The objective function was to find the ball mix that guarantees maximum production of the floatable size range (i.e. 75+9m) applicable to a platinum-bearing ore.
The simulated product size distributions were found to display a close match with the measured product from an operational mill. Based on the maximum production of floatable particles, the best combination of make-up ball size was determined in relation to feed flow rate, feed size distribution and ball filling.
Also called silica sand or quartz sand, silica is silicon dioxide (SiO2). Silicon compounds are the most significant component of the Earths crust. Since sand is plentiful, easy to mine and relatively easy to process, it is the primary ore source of silicon. The metamorphic rock, quartzite, is another source.
Silicon (Si) is a semi-metallic or metalloid, because it has several of the metallic characteristics. Silicon is never found in its natural state, but rather in combination with oxygen as the silicate ion SiO44-in silica-rich rocks such as obsidian, granite, diorite, and sandstone. Feldspar and quartz are the most significant silicate minerals. Silicon alloys include a variety of metals, including iron, aluminum, copper, nickel, manganese and ferrochromium.
In almost all cases, silica mining uses open pit or dredging mining methods with standard mining equipment. Except for temporarily disturbing the immediate area while mining operations are active, sand and gravel mining usually has limited environmental impact.
In addition to tool steels, an example of alloy steels, ferrosilicon is used in the manufacture of stainless steels, carbon steels, and other alloy steels. An alloy steel refers to all finished steels other than stainless and carbon steels. Stainless steels are used when superior corrosion resistance, hygiene, aesthetic, and wear-resistance qualities are needed.
Silicon is used in the aluminum industry to improve castability and weldability. Silicon-aluminum alloys tend to have relatively low strength and ductility, so other metals, especially magnesium and copper, are often added to improve strength.
In the chemicals industry, silicon metal is the starting point for the production of silianes, silicones, fumed silica, and semiconductor-grade silicon. Silanes are the used to make silicone resins, lubricants, anti-foaming agents, and water-repellent compounds. Silicones are used as lubricants, hydraulic fluids, electrical insulators, and moisture-proof treatments.
Semiconductor-grade silicon is used in the manufacture of silicon chips and solar cells. Fumed silica is used as a filler in the cement and refractory materials industries, as well as in heat insulation and filling material for synthetic rubbers, polymers and grouts.
Silicon is considered a semiconductor. This means that it conducts electricity, but not as well as a metal such as copper or silver. This physical property makes silicon an important commodity in the computer manufacturing business.
Foundry silica sand is the quartz as the main mineral composition, particle size of 0.020 mm to 3.350 mm refractory particles, according to the mining and processing methods of different can be divided into artificial silica sand and sand washing, sand washing, selection of natural silica sand such as sand (flotation).
1, crushing process directly.Its technological process is: the run of mine ore grizzly jaw crusher crushing and screening to cone crushing and screening, more paragraphs to roll the crushing and screening products.
3, since the mill grinding process.Its technological process is: the run of mine ore grizzly since the mill to air classification system, qualified products and super fine powder (coarse grained grinding cage hexagon screen the final product).
4, wet rod mill magnetic separation process.This is China during the period of > study of success and is promoting the new technology, its technological process is: in ore to coarse crushing, crushing and screening to rod mill, the high frequency fine screen to hydraulic classification, magnetic separation and iron products.
Material: silica sand Capacity: 100TPH Country: Malaysia Feeding size: 6mm Raw mineral description: 1. Sand particle distribution: <1mm 40%; 1-3mm 30%; 3-6mm 30% 2. Mineral composition: silica, SiO2 94.7%; Iron as Fe2O3 0.35%; Alumina, Al2O3 0.78%Customers requirements: Silica concentrate with >98.5% of SiO2.
Project: extraction of scrubbed silica sand (use for glass, cement industry) Material: silica sand Capacity: 50TPH Country: Indonesia Mineral condition description: raw silica sand contains slime and barren rock, which is up to 25-30 CM in diameter. Customers requirements: wash the mud off and sieve the barren rock out.
Material: silica sand Capacity: 65TPH Country: Malaysia Feeding size: 0-1mm Raw mineral description: 1. Mineral composition: silicon, SiO2 94%; Iron as Fe2O3 0.35%; others 2. Customer roughly washed the sand and sieved 0-1mm fine sand out.
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