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In all ore dressing and milling Operations, including flotation, cyanidation, gravity concentration, and amalgamation, the Working Principle is to crush and grind, often with rob mill & ball mills, the ore in order to liberate the minerals. In the chemical and process industries, grinding is an important step in preparing raw materials for subsequent treatment.In present day practice, ore is reduced to a size many times finer than can be obtained with crushers. Over a period of many years various fine grinding machines have been developed and used, but the ball mill has become standard due to its simplicity and low operating cost.
A ball millefficiently operated performs a wide variety of services. In small milling plants, where simplicity is most essential, it is not economical to use more than single stage crushing, because the Steel-Head Ball or Rod Mill will take up to 2 feed and grind it to the desired fineness. In larger plants where several stages of coarse and fine crushing are used, it is customary to crush from 1/2 to as fine as 8 mesh.
Many grinding circuits necessitate regrinding of concentrates or middling products to extremely fine sizes to liberate the closely associated minerals from each other. In these cases, the feed to the ball mill may be from 10 to 100 mesh or even finer.
Where the finished product does not have to be uniform, a ball mill may be operated in open circuit, but where the finished product must be uniform it is essential that the grinding mill be used in closed circuit with a screen, if a coarse product is desired, and with a classifier if a fine product is required. In most cases it is desirable to operate the grinding mill in closed circuit with a screen or classifier as higher efficiency and capacity are obtained. Often a mill using steel rods as the grinding medium is recommended, where the product must have the minimum amount of fines (rods give a more nearly uniform product).
Often a problem requires some study to determine the economic fineness to which a product can or should be ground. In this case the 911Equipment Company offers its complete testing service so that accurate grinding mill size may be determined.
Until recently many operators have believed that one particular type of grinding mill had greater efficiency and resulting capacity than some other type. However, it is now commonly agreed and accepted that the work done by any ballmill depends directly upon the power input; the maximum power input into any ball or rod mill depends upon weight of grinding charge, mill speed, and liner design.
The apparent difference in capacities between grinding mills (listed as being the same size) is due to the fact that there is no uniform method of designating the size of a mill, for example: a 5 x 5 Ball Mill has a working diameter of 5 inside the liners and has 20 per cent more capacity than all other ball mills designated as 5 x 5 where the shell is 5 inside diameter and the working diameter is only 48 with the liners in place.
Ball-Rod Mills, based on 4 liners and capacity varying as 2.6 power of mill diameter, on the 5 size give 20 per cent increased capacity; on the 4 size, 25 per cent; and on the 3 size, 28 per cent. This fact should be carefully kept in mind when determining the capacity of a Steel- Head Ball-Rod Mill, as this unit can carry a greater ball or rod charge and has potentially higher capacity in a given size when the full ball or rod charge is carried.
A mill shorter in length may be used if the grinding problem indicates a definite power input. This allows the alternative of greater capacity at a later date or a considerable saving in first cost with a shorter mill, if reserve capacity is not desired. The capacities of Ball-Rod Mills are considerably higher than many other types because the diameters are measured inside the liners.
The correct grinding mill depends so much upon the particular ore being treated and the product desired, that a mill must have maximum flexibility in length, type of grinding medium, type of discharge, and speed.With the Ball-Rod Mill it is possible to build this unit in exact accordance with your requirements, as illustrated.
To best serve your needs, the Trunnion can be furnished with small (standard), medium, or large diameter opening for each type of discharge. The sketch shows diagrammatic arrangements of the four different types of discharge for each size of trunnion opening, and peripheral discharge is described later.
Ball-Rod Mills of the grate discharge type are made by adding the improved type of grates to a standard Ball-Rod Mill. These grates are bolted to the discharge head in much the same manner as the standard headliners.
The grates are of alloy steel and are cast integral with the lifter bars which are essential to the efficient operation of this type of ball or rod mill. These lifter bars have a similar action to a pump:i. e., in lifting the product so as to discharge quickly through the mill trunnion.
These Discharge Grates also incorporate as an integral part, a liner between the lifters and steel head of the ball mill to prevent wear of the mill head. By combining these parts into a single casting, repairs and maintenance are greatly simplified. The center of the grate discharge end of this mill is open to permit adding of balls or for adding water to the mill through the discharge end.
Instead of being constructed of bars cast into a frame, Grates are cast entire and have cored holes which widen toward the outside of the mill similar to the taper in grizzly bars. The grate type discharge is illustrated.
The peripheral discharge type of Ball-Rod Mill is a modification of the grate type, and is recommended where a free gravity discharge is desired. It is particularly applicable when production of too many fine particles is detrimental and a quick pass through the mill is desired, and for dry grinding.
The drawings show the arrangement of the peripheral discharge. The discharge consists of openings in the shell into which bushings with holes of the desired size are inserted. On the outside of the mill, flanges are used to attach a stationary discharge hopper to prevent pulp splash or too much dust.
The mill may be operated either as a peripheral discharge or a combination or peripheral and trunnion discharge unit, depending on the desired operating conditions. If at any time the peripheral discharge is undesirable, plugs inserted into the bushings will convert the mill to a trunnion discharge type mill.
Unless otherwise specified, a hard iron liner is furnished. This liner is made of the best grade white iron and is most serviceable for the smaller size mills where large balls are not used. Hard iron liners have a much lower first cost.
Electric steel, although more expensive than hard iron, has advantage of minimum breakage and allows final wear to thinner section. Steel liners are recommended when the mills are for export or where the source of liner replacement is at a considerable distance.
Molychrome steel has longer wearing qualities and greater strength than hard iron. Breakage is not so apt to occur during shipment, and any size ball can be charged into a mill equipped with molychrome liners.
Manganese liners for Ball-Rod Mills are the world famous AMSCO Brand, and are the best obtainable. The first cost is the highest, but in most cases the cost per ton of ore ground is the lowest. These liners contain 12 to 14% manganese.
The feed and discharge trunnions are provided with cast iron or white iron throat liners. As these parts are not subjected to impact and must only withstand abrasion, alloys are not commonly used but can be supplied.
Gears for Ball-Rod Mills drives are furnished as standard on the discharge end of the mill where they are out of the way of the classifier return, scoop feeder, or original feed. Due to convertible type construction the mills can be furnished with gears on the feed end. Gear drives are available in two alternative combinations, which are:
All pinions are properly bored, key-seated, and pressed onto the steel countershaft, which is oversize and properly keyseated for the pinion and drive pulleys or sheaves. The countershaft operates on high grade, heavy duty, nickel babbitt bearings.
Any type of drive can be furnished for Ball-Rod Mills in accordance with your requirements. Belt drives are available with pulleys either plain or equipped with friction clutch. Various V- Rope combinations can also be supplied.
The most economical drive to use up to 50 H. P., is a high starting torque motor connected to the pinion shaft by means of a flat or V-Rope drive. For larger size motors the wound rotor (slip ring) is recommended due to its low current requirement in starting up the ball mill.
Should you be operating your own power plant or have D. C. current, please specify so that there will be no confusion as to motor characteristics. If switches are to be supplied, exact voltage to be used should be given.
Even though many ores require fine grinding for maximum recovery, most ores liberate a large percentage of the minerals during the first pass through the grinding unit. Thus, if the free minerals can be immediately removed from the ball mill classifier circuit, there is little chance for overgrinding.
This is actually what has happened wherever Mineral Jigs or Unit Flotation Cells have been installed in the ball mill classifier circuit. With the installation of one or both of these machines between the ball mill and classifier, as high as 70 per cent of the free gold and sulphide minerals can be immediately removed, thus reducing grinding costs and improving over-all recovery. The advantage of this method lies in the fact that heavy and usually valuable minerals, which otherwise would be ground finer because of their faster settling in the classifier and consequent return to the grinding mill, are removed from the circuit as soon as freed. This applies particularly to gold and lead ores.
Ball-Rod Mills have heavy rolled steel plate shells which are arc welded inside and outside to the steel heads or to rolled steel flanges, depending upon the type of mill. The double welding not only gives increased structural strength, but eliminates any possibility of leakage.
Where a single or double flanged shell is used, the faces are accurately machined and drilled to template to insure perfect fit and alignment with the holes in the head. These flanges are machined with male and female joints which take the shearing stresses off the bolts.
The Ball-Rod Mill Heads are oversize in section, heavily ribbed and are cast from electric furnace steel which has a strength of approximately four times that of cast iron. The head and trunnion bearings are designed to support a mill with length double its diameter. This extra strength, besides eliminating the possibility of head breakage or other structural failure (either while in transit or while in service), imparts to Ball-Rod Mills a flexibility heretofore lacking in grinding mills. Also, for instance, if you have a 5 x 5 mill, you can add another 5 shell length and thus get double the original capacity; or any length required up to a maximum of 12 total length.
On Type A mills the steel heads are double welded to the rolled steel shell. On type B and other flanged type mills the heads are machined with male and female joints to match the shell flanges, thus taking the shearing stresses from the heavy machine bolts which connect the shell flanges to the heads.
The manhole cover is protected from wear by heavy liners. An extended lip is provided for loosening the door with a crow-bar, and lifting handles are also provided. The manhole door is furnished with suitable gaskets to prevent leakage.
The mill trunnions are carried on heavy babbitt bearings which provide ample surface to insure low bearing pressure. If at any time the normal length is doubled to obtain increased capacity, these large trunnion bearings will easily support the additional load. Trunnion bearings are of the rigid type, as the perfect alignment of the trunnion surface on Ball-Rod Mills eliminates any need for the more expensive self-aligning type of bearing.
The cap on the upper half of the trunnion bearing is provided with a shroud which extends over the drip flange of the trunnion and effectively prevents the entrance of dirt or grit. The bearing has a large space for wool waste and lubricant and this is easily accessible through a large opening which is covered to prevent dirt from getting into the bearing.Ball and socket bearings can be furnished.
Scoop Feeders for Ball-Rod Mills are made in various radius sizes. Standard scoops are made of cast iron and for the 3 size a 13 or 19 feeder is supplied, for the 4 size a 30 or 36, for the 5 a 36 or 42, and for the 6 a 42 or 48 feeder. Welded steel scoop feeders can, however, be supplied in any radius.
The correct size of feeder depends upon the size of the classifier, and the smallest feeder should be used which will permit gravity flow for closed circuit grinding between classifier and the ball or rod mill. All feeders are built with a removable wearing lip which can be easily replaced and are designed to give minimum scoop wear.
A combination drum and scoop feeder can be supplied if necessary. This feeder is made of heavy steel plate and strongly welded. These drum-scoop feeders are available in the same sizes as the cast iron feeders but can be built in any radius. Scoop liners can be furnished.
The trunnions on Ball-Rod Mills are flanged and carefully machined so that scoops are held in place by large machine bolts and not cap screws or stud bolts. The feed trunnion flange is machined with a shoulder for insuring a proper fit for the feed scoop, and the weight of the scoop is carried on this shoulder so that all strain is removed from the bolts which hold the scoop.
High carbon steel rods are recommended, hot rolled, hot sawed or sheared, to a length of 2 less than actual length of mill taken inside the liners. The initial rod charge is generally a mixture ranging from 1.5 to 3 in diameter. During operation, rod make-up is generally the maximum size. The weights per lineal foot of rods of various diameters are approximately: 1.5 to 6 lbs.; 2-10.7 lbs.; 2.5-16.7 lbs.; and 3-24 lbs.
Forged from the best high carbon manganese steel, they are of the finest quality which can be produced and give long, satisfactory service. Data on ball charges for Ball-Rod Mills are listed in Table 5. Further information regarding grinding balls is included in Table 6.
Rod Mills has a very define and narrow discharge product size range. Feeding a Rod Mill finer rocks will greatly impact its tonnage while not significantly affect its discharge product sizes. The 3.5 diameter rod of a mill, can only grind so fine.
Crushers are well understood by most. Rod and Ball Mills not so much however as their size reduction actions are hidden in the tube (mill). As for Rod Mills, the image above best expresses what is going on inside. As rocks is feed into the mill, they are crushed (pinched) by the weight of its 3.5 x 16 rods at one end while the smaller particles migrate towards the discharge end and get slightly abraded (as in a Ball Mill) on the way there.
We haveSmall Ball Mills for sale coming in at very good prices. These ball mills are relatively small, bearing mounted on a steel frame. All ball mills are sold with motor, gears, steel liners and optional grinding media charge/load.
Ball Mills or Rod Mills in a complete range of sizes up to 10 diameter x20 long, offer features of operation and convertibility to meet your exactneeds. They may be used for pulverizing and either wet or dry grindingsystems. Mills are available in both light-duty and heavy-duty constructionto meet your specific requirements.
All Mills feature electric cast steel heads and heavy rolled steelplate shells. Self-aligning main trunnion bearings on large mills are sealedand internally flood-lubricated. Replaceable mill trunnions. Pinion shaftbearings are self-aligning, roller bearing type, enclosed in dust-tightcarrier. Adjustable, single-unit soleplate under trunnion and drive pinionsfor perfect, permanent gear alignment.
Ball Mills can be supplied with either ceramic or rubber linings for wet or dry grinding, for continuous or batch type operation, in sizes from 15 x 21 to 8 x 12. High density ceramic linings of uniform hardness male possible thinner linings and greater and more effective grinding volume. Mills are shipped with liners installed.
Complete laboratory testing service, mill and air classifier engineering and proven equipment make possible a single source for your complete dry-grinding mill installation. Units available with air swept design and centrifugal classifiers or with elevators and mechanical type air classifiers. All sizes and capacities of units. Laboratory-size air classifier also available.
A special purpose batch mill designed especially for grinding and mixing involving acids and corrosive materials. No corners mean easy cleaning and choice of rubber or ceramic linings make it corrosion resistant. Shape of mill and ball segregation gives preferential grinding action for grinding and mixing of pigments and catalysts. Made in 2, 3 and 4 diameter grinding drums.
Nowadays grinding mills are almost extensively used for comminution of materials ranging from 5 mm to 40 mm (3/161 5/8) down to varying product sizes. They have vast applications within different branches of industry such as for example the ore dressing, cement, lime, porcelain and chemical industries and can be designed for continuous as well as batch grinding.
Ball mills can be used for coarse grinding as described for the rod mill. They will, however, in that application produce more fines and tramp oversize and will in any case necessitate installation of effective classification.If finer grinding is wanted two or three stage grinding is advisable as for instant primary rod mill with 75100 mm (34) rods, secondary ball mill with 2540 mm(11) balls and possibly tertiary ball mill with 20 mm () balls or cylpebs.To obtain a close size distribution in the fine range the specific surface of the grinding media should be as high as possible. Thus as small balls as possible should be used in each stage.
The principal field of rod mill usage is the preparation of products in the 5 mm0.4 mm (4 mesh to 35 mesh) range. It may sometimes be recommended also for finer grinding. Within these limits a rod mill is usually superior to and more efficient than a ball mill. The basic principle for rod grinding is reduction by line contact between rods extending the full length of the mill, resulting in selective grinding carried out on the largest particle sizes. This results in a minimum production of extreme fines or slimes and more effective grinding work as compared with a ball mill. One stage rod mill grinding is therefore suitable for preparation of feed to gravimetric ore dressing methods, certain flotation processes with slime problems and magnetic cobbing. Rod mills are frequently used as primary mills to produce suitable feed to the second grinding stage. Rod mills have usually a length/diameter ratio of at least 1.4.
Tube mills are in principle to be considered as ball mills, the basic difference being that the length/diameter ratio is greater (35). They are commonly used for surface cleaning or scrubbing action and fine grinding in open circuit.
In some cases it is suitable to use screened fractions of the material as grinding media. Such mills are usually called pebble mills, but the working principle is the same as for ball mills. As the power input is approximately directly proportional to the volume weight of the grinding media, the power input for pebble mills is correspondingly smaller than for a ball mill.
A dry process requires usually dry grinding. If the feed is wet and sticky, it is often necessary to lower the moisture content below 1 %. Grinding in front of wet processes can be done wet or dry. In dry grinding the energy consumption is higher, but the wear of linings and charge is less than for wet grinding, especially when treating highly abrasive and corrosive material. When comparing the economy of wet and dry grinding, the different costs for the entire process must be considered.
An increase in the mill speed will give a directly proportional increase in mill power but there seems to be a square proportional increase in the wear. Rod mills generally operate within the range of 6075 % of critical speed in order to avoid excessive wear and tangled rods. Ball and pebble mills are usually operated at 7085 % of critical speed. For dry grinding the speed is usually somewhat lower.
The mill lining can be made of rubber or different types of steel (manganese or Ni-hard) with liner types according to the customers requirements. For special applications we can also supply porcelain, basalt and other linings.
The mill power is approximately directly proportional to the charge volume within the normal range. When calculating a mill 40 % charge volume is generally used. In pebble and ball mills quite often charge volumes close to 50 % are used. In a pebble mill the pebble consumption ranges from 315 % and the charge has to be controlled automatically to maintain uniform power consumption.
In all cases the net energy consumption per ton (kWh/ton) must be known either from previous experience or laboratory tests before mill size can be determined. The required mill net power P kW ( = ton/hX kWh/ton) is obtained from
Trunnions of S.G. iron or steel castings with machined flange and bearing seat incl. device for dismantling the bearings. For smaller mills the heads and trunnions are sometimes made in grey cast iron.
The mills can be used either for dry or wet, rod or ball grinding. By using a separate attachment the discharge end can be changed so that the mills can be used for peripheral instead of overflow discharge.
This Grindability Test orBond Ball Mill Work Index Procedure is used to determine the Bond Work Index of minus six mesh or finer feed ore samples. These equation application methods are used to process <1/2 ore samples in a Ball Mill using a standard ball charge.
Below describes in general terms the Bond Work Index Procedure used by a Professional Metallurgical Testing Laboratory. If you think the equation is complicated, wait until you try using that famous Bond Work Index equation Mr. Fred worked so hard at.
9. Enter the combined sizing data from Feed No. 1 and No. 2 (or No. 2 and No. 3) into a new Bond Feed Screen Analysis sheet and calculate the K80 and the average percent of less than 150 mesh product. 10. Record the K80 in um as factor, F, on yourBond Mill Grindability Test-sheet. 11. Record the percent of the under-size, i.e. less than 150 mesh product.
The sample was crushed to 100% passing 6 mesh (3.35mm), from this a 700 cc volume was measured and weighed to be used as feed for the Bond Mill. A feed particle size analysis was performed to determine feed P80 and % -106 micron. The sample was milled for 100 revolutions and passed over a 150 mesh (106m) sieve, the undersize was removed from the samples, weighed, and new sample was added to the feed to maintain the initial sample weights. The number of rotations for the next milling was calculated based on the undersize sample weights. This was repeated until the milling produced stable results for three consecutive cycles.
The Bond grindability was calculated to be 1.21 g net undersize/revolution by averaging the results from the last 3 runs on the Bond Mill. The grindability is defined as the ease at which a mineral particle is reduced to a predetermined size and used to calculate the Work Index. The feed and final undersize particle size analysis as well as the grindability test data sheet are found in the appendix.
Portland cement clinker is nodules (diameters, 525mm) of sintered material produced by heating a homogeneous mixture of raw materials in a kiln to a sintering temperature of approximately 1450C for modern cements.
Portland cement is a fine powder produced by grinding Portland cement clinker (more than 90%), a limited amount of gypsum (calcium sulphate dehydrate CaSO4.2H2O, which controls the set time) and other minor constituents which can be used to vary the properties of the final cement. The standard Portland cement is often referred to as Ordinary Portland Cement, and European Standard EN197-1 gives the following description:
Portland cement clinker is a hydraulic material which shall consist of at least two-thirds by mass of calcium silicates (3CaO.SiO2 and 2CaO.SiO2), the remainder consisting of aluminium and iron containing clinker phases and other compounds. The ratio of CaO to SiO2 shall not be less than 2.0. The magnesium oxide content (MgO) shall not exceed 5.0% by mass.
Hydraulic cement (cement that not only hardens by reacting with water but also forms a water-resistant product) produced by pulverising clinkers consisting essentially of hydraulic calcium silicates, usually containing one or more of the forms of calcium sulphate as an inter-ground addition.
Portland cement clinker is nodules (diameters, 525mm) of sintered material produced by heating a homogeneous mixture of raw materials in a kiln to a sintering temperature of approximately 1450C for modern cements. The resulting clinker consists of four main minerals:11
The aluminium oxide and iron oxide are present in the main as a flux and contribute little to the mechanical strength of the final concrete. The proportions of each mineral in the clinker are important in determining the properties of the resulting cement. For example, in some special cements, such as Low Heat (LH) and Sulphate Resistant (SR) types, it is necessary to limit the amount of tricalcium aluminate (3CaO-Al2O3) that is formed.
Portland cement clinker contains four principal chemical compounds, which are normally referred to as the clinker minerals. The composition of the minerals and their normal range of levels in current UK and European Portland cement clinkers are summarized in Table1.1.
It is the two calcium silicate minerals, C3S and C2S, which are largely responsible for the strength development and the long-term structural and durability properties of Portland cement. However, the reaction between CaO (lime from limestone) and SiO2 (silica from sand) is very difficult to achieve, even at high firing temperatures. Chemical combination is greatly facilitated if small quantities of alumina and iron oxide are present (typically 5% Al2O3 and 3% Fe2O3), as these help to form a molten flux through which the lime and silica are able to partially dissolve, and then react to yield C3S and C2S. The sequence of reactions, which take place in the kiln, is illustrated in Figure1.5.
The reaction requiring the greatest energy input is the decarbonation of CaCO3, which takes place mainly in the temperature range 7001000C. For a typical mix containing 80% limestone the energy input to decarbonate the CaCO3 is approximately 400 kCal/kg of clinker, which is approximately half of the total energy requirement of a modem dry process kiln.
When decarbonation is complete at about 1100C, the feed temperature rises more rapidly. Lime reacts with silica to form belite (C2S) but the level of unreacted lime remains high until a temperature of ~1250C is reached. This is the lower limit of thermodynamic stability of alite (C3S). At ~1300C partial melting occurs, the liquid phase (or flux) being provided by the alumina and iron oxide present. The level of unreacted lime reduces as C2S is converted to C3S. The process will be operated to ensure that the level of unreacted lime (free lime) is below 3%.
Normally, C3S formation is effectively complete at a material temperature of about 1450C, and the level of uncombined lime reduces only slowly with further residence time. The ease with which the clinker can be combined is strongly influenced by the mineralogy of the raw materials and, in particular, the level of coarse silica (quartz) present. The higher the level of coarse silica in the raw materials, the finer the raw mix will have to be ground to ensure satisfactory combination at acceptable kiln temperatures.
Coarse silica is also associated with the occurrence of clusters of relatively large belite crystals around the sites of the silica particles. Figures1.6(a) and 1.6(b) are photomicrographs of a normal clinker containing well-distributed alite and belite and clinker produced from a raw meal containing relatively coarse silica.
As the clinker passes under the flame it starts to cool and the molten C3A and C4AF, which constitute the flux phase, crystallize. This crystallization is normally complete by the time the clinker exits the rotary kiln and enters the cooler at a temperature of ~1200C. Slow cooling should be avoided as this can result in an increase in the belite content at the expense of alite and also the formation of relatively large C3A crystals which can result in unsatisfactory concrete rheology (water demand and stiffening).
Portland cement clinker, as ordinarily prepared, is not a homogeneous substance, but a rather fine-grained mixture of several solid phases. It is therefore difficult to draw any conclusions from a study of its chemical reactions alone, since these may involve more than one constituent, and the only methods capable of yielding trustworthy results are those which enable us to deal with the individual constituents separately. The conditions have a close parallel with the study of igneous rocks. It would be impossible to determine the structure of a granite by observing its gross behaviour towards reagents. The reactions with the quartz, feldspar and mica would be superimposed and confused, and the resulting action would give only a meaningless average. A number of methods are available for overcoming these difficulties, such as mineral separation after crushing, optical examination of the crushed material or of thin section or of polished and etched surfaces, the use of X-rays, and electron microprobe analysis.
A rock may be crushed to such a degree of fineness as to release the constituent crystal grains and the powder thus obtained can be suspended in liquids of suitable densities, bringing about a separation of the light and heavier minerals. In skilled hands, the method is capable of giving very accurate results. It has been applied to cements, but with less success, as the close intermixture and friable character of the constituents render a separation impossible until the whole has been reduced to a fine powder, when the particles no longer settle satisfactorily after suspension in heavy liquids. An improvement may be obtained by effecting the separation of the crushed material by centrifugal or magnetic means, but even then a clean separation has not been effected. Microscopic examination of the crushed material shows that many of the grains are composed of one constituent, but that attached to their edges there may remain fragments of a second constituent, thus rendering a perfect separation impracticable. Such partially successful separations, using chemical or physical methods, can be a useful preliminary to X-ray examination since the intensity of the reflections obtained for a given constituent will be enhanced.
Tricalcium silicate grows to a larger size than any other constituent, and separation of enough material for analysis by centrifuging a graded clinker powder in heavy liquids has been achieved. Partial hydration35 of a clinker left the slowly hydrating 2CaOSiO2 as a residue after acid extraction of the set cement. Many workers have used an alkaline solution of a dimethylamine salt to dissolve the silicate phases, leaving a concentration of other constituents. A solution of salicyclic acid in methanol has also been used to dissolve the calcium silicate phases from Portland cement.4649 The cement specimen must first be ground to a particle size less than 5 m in an agate ball mill, with cyclohexane as a grinding aid. Other organic acid solutions have been proposed.46 These methods, or alkaline ammonium citrate or acetic acid,5052 have been used to intensify the X-ray reflections from the ferrite phase in order to obtain its composition from the d-spacings. A solution of KOH containing sucrose is recommended as a dissolution medium which removes the calcium aluminate phases.49 These solubility methods depend on differences in the rate of solution of the various compounds and require, therefore, a suitable choice of the extraction conditions. A bar magnet has also been used to separate the ferrite phase.53
The sintering of Portland cement clinker is simply called twice grinding and once sintering, that is, grinding the cement with cement raw material; sintering the calcined part of the raw material into clinker; grinding the clinker with a limited amount of gypsum into Portland cement clinker. The sintering process of Portland cement clinker is shown in Figure4.1.
Different proportions of cement raw materials directly affect the proportions of the mineral components of Portland cement clinker and the main building technical performance. The process that cement raw materials is sintered in a kiln is the key to the quality of cement clinker.
In the sintering process of cement raw material, the useful components decomposed by various raw materials at 1000C are mainly: calcium oxide (CaO), silicon dioxide (SiO2), aluminum oxide (Al2O3), and ferric oxide (Fe2O3). And the solid-state reaction occurs to a small amount of oxide at about 800C which generates calcium aluminate, a small amount of dicalcium ferrite, and dicalcium silicate.
At 1300~1450C, tricalcium aluminate and tetracalcium aluminoferrite are in molten state and CaO and part of dicalcium silicate are dissolved in the generated liquid phase. In this liquid phase, dicalcium silicate synthesizes tricalcium silicate by absorbing CaO which is the key to the sintering of cement. Sufficient time should be cost to make the free CaO in the raw material be absorbed, for the quality of cement clinker.
Once the Portland cement clinker has been manufactured it is normally fed to a store to effect a measure of blending and also to allow it to cool to ambient temperature. The latter operation is desirable because most of the clinker coolers associated with kiln operation are unable to lower the temperature below 5080C, and even at this temperature the amount of heat introduced into the grinding process is unwelcome.
At most manufacturing plants the ball mill is used to grind the clinker and, since the production rate is directly related to the amount of electrical energy supplied, the power of the electric motor used to turn the mill is a first-order measure of the output achieved. Mills vary in their power input from as little as 200 kW up to 10000 kW.
An efficient mill system grinding a Portland cement (without secondary components) to a level of fineness required for a 42.5 strength class4 can be expected to consume of the order of 30 kWh/t, and on this basis the mills cited above should be capable of producing 67 and 333 t/h, respectively.
The ball mill in its simplest form consists of a tube rotating about a horizontal axis. The inside is normally divided into at least two chambers separated by slotted diaphragm(s). This division enables the mill to operate with at least two different size gradings of grinding media (usually balls). Such an arrangement is necessary because the clinker normally fed to the mill to be ground can contain lumps as large as 60 mm in diameter. These large lumps require a larger ball size (90100 mm) to break them down. At the same time, balls as small as 13 mm in diameter are needed to grind the material to the fineness required in the more rapid-hardening cements. If such a wide range of ball sizes were to be placed into one chamber, the smaller balls would move to the inlet end and the larger balls to the outlet end, thus making effective grinding impossible. By segregating the ball sizes through the use of diaphragms the larger balls can be kept at the inlet end of the mill and the smaller balls at the outlet. It is possible to achieve a similar effect in a single chamber through the use of classifying liner plates in the mill. These are plates with a wedge-type profile with the tapered part of the wedge facing the mill inlet. These are effective in dealing with ball sizes within the range 6013 mm, and it is normal practice where larger clinker particle sizes have to be dealt with to introduce a separate chamber accommodating 9060 mm diameter media.
In order to achieve efficient grinding it is necessary for the media to cascade over each other as the mill rotates. This means that the rotational speed must be kept below the critical point where the media are held against the mill shell by centrifugal force. This critical speed is defined as 42.3 divided by the square root of the internal diameter of the mill (in metres), and the rotational speed of most mills is kept within 6580 per cent of the critical speed.
The volume of grinding media used in the mill is normally established at a level which either gives the lowest specific power consumption (kWh/t of product) or gives the maximum output. In the case of the latter, it is often decided by the maximum amount of power available from the motor. However, subject to the motor power available, a good starting point would be 30 per cent of the internal volume of the mill taken up with the grinding media. A useful relationship between mill power, media loading speed and rotational speed is given by the following formula:
where D = internal diameter of the mill (m); A = a constant for a given mill system, usually about 0.245; W = the mass of the grinding media in tonnes; and N = the rotational speed of the mill (rev/min).
Only 12 per cent of the electrical energy supplied to the ball mill is used in actually fracturing the particles. This means that in the course of the grinding operation a not inconsiderable amount of heat is produced, and one of the principle problems associated with cement grinding is to remove the heat. In the case of small output mills (up to 900 kW) fed with cold clinker it is possible to achieve an adequate degree of cooling by spraying water on to the mill shell. However, with larger mills the ratio of shell surface to heat input decreases and other methods have to be adopted. One of the most satisfactory is to spray water into the outlet chamber of the mill and to use the latent heat of vaporisation to remove the heat. This requires good control facilities and also sufficient ventilation to prevent the water hydrating the cement and causing a loss in its strength-giving properties. Other methods which are used, often in conjunction with water injection, are to introduce a classifier into the milling system. This involves using the mill to produce a relatively coarse product (having, say, a Blaine specific surface area of 270 m2/kg) and to use the classifier to separate out a product of the required fineness. The power required to produce a coarse product is less and hence the heat introduced is also less. In addition, the movement of the hot cement in an elevator and through the classifier, particularly if it is air swept, can effect a considerable degree of cooling. In this type of mill system -known as a closed-circuit mill the coarse material produced in the classifier is fed back to the feed end of the ball mill. Closed-circuit grinding is necessary when large mills are required to produce finely ground cements and also in situations where there is a significant difference in the hardness of the Portland cement clinker and any secondary material incorporated into the cement. Grinding cement in a closed-circuit mill system generally produces a narrower particle grading in the cement than that produced from a mill system without a classifier (an open-circuit mill) and this in turn leads to a higher water demand in the mortar or concrete produced from that cement. The higher water demand means that, for a given cement content, lower strengths are produced.
The emphasis placed upon grinding temperatures is associated with the effect that this has upon the calcium sulfate (gypsum rock) added to the clinker, in amounts normally between 3 and 8 per cent, to retard the hydration of the tricalcium aluminate and to optimise the strength-giving properties of the calcium silicates.
If the calcium sulfate is added in the form of gypsum, it can become dehydrated at grinding temperatures of the order of 115130C or above, and can be present in a form which when mixed with water forms a supersaturated solution with respect to gypsum and from which secondary gypsum precipitates to provide a structure having some rigidity. This has the effect of stiffening the concrete or mortar and making it necessary to add additional water and thereby lower the strength-giving properties. This is known as false set and should be distinguished from flash set, which results in cements which have insufficient sulfate present effectively to stop the hydration of the tricalcium aluminate to the hydrate rather than to ettringite. Flash set is accompanied with the release of considerable amounts of heat whereas it is sometimes possible with false set (as its name implies) to break down the structure developed through vigorous mixing. Some of the problems with false setting can be alleviated by replacing some of the gypsum added to the cement by natural anhydrite. Other problems associated with the calcium sulfate addition arise when the clinker contains a relatively high SO5 content. As most cement specifications contain a requirement for a maximum SO3 level, the amount of calcium sulfate which can be added is restricted, and in that situation must be in a form where the solubility is sufficient to prevent the formation of the tricalcium aluminate hydrate rather than ettringite.
Other problems occur when the reactivity of the tricalcium aluminate is reduced through the uptake of moisture and/or carbon dioxide.107 In this case it may be unable to react sufficiently rapidly to remove enough sulfate from the aqueous phase to prevent the precipitation of secondary gypsum, although when in its original (unaerated) state this situation prevails.
The pre-eminence of the ball mill has been challenged by the vertical spindle (or roller) mill on the grounds that it is capable of a lower power consumption per unit mass of product (kWh/t). However, the particle grading produced from such a mill tends to be narrower than the ball mill and products made in this way tend to suffer from relatively high water demands when made into concrete or mortar. More work on this approach would appear to be required before any benefits can be realised.
Another development is the introduction of the roller press (otherwise known as high-pressure material-bed comminution)108 to disintegrate the clinker. This involves the use of two rollers turning at a peripheral speed of between 0.9 and 1.8 m/s and with a gap between them of 830 mm. The pressure developed on the particles exceeds 50 MPa, and claims have been made that the clinker is activated by the very intensive stressing which occurs as a result of the passage through the rollers. It has been suggested that, provided the product from the roller press is operated in closed circuit with a disintegrator and a classifier, a ball mill may become unnecessary. If this is done, power savings of 45 per cent have been claimed.109 However, despite this, the normal application of the roller press is in conjunction with a ball mill, taking advantage of the smaller feedstock size to use it in the single chamber configuration. Power savings of 30 per cent have been claimed.
In the alkali-free samples, setting time is reduced slightly with increasing silica ratio and with increasing alumina ratio. The setting times of cements made from alkali and sulfate free clinkers were reduced with increasing lime saturation factor.
This formula represents fairly well the maximum amount of lime that can be combined in Portland cement clinker. A refinement for the maximum combined lime content to allow for the small amount of MgO combined in 3CaOSiO2 has been suggested31:
The maximum value for % MgO that can be inserted in this formula is ~2%, since any excess tends to be present as free MgO (as the mineral periclase) after firing, thus rendering the formula inappropriate.
In the manufacturing of PC clinker, the raw materials are mixed and heated to temperatures up to 1450C. To identify the potential phases after heating the raw mix blend, the lime saturation factor (LSF) is often used to verify the ratio of C3S to C2S. It also shows whether the clinker is likely to contain an unacceptable proportion of free lime. Values between 0.92 and 0.98 are typical of modern clinkers, and a mix with an LSF greater than 1.0 will yield free CaO, which is liable to persist in the final product, regardless of the degree of mixing and time during which the clinkering temperature is maintained (Taylor, 1997).
The silica ratio and alumina ratio (also respectively called silica modulus and alumina modulus) are empirically used to characterise the potential mineralogical composition of the cement clinker. The silica modulus mainly governs the proportion of silicate phases in the clinker, whilst the alumina modulus governs the ratio of aluminate to ferrite phases in the clinker; for normal PC clinker, the silica and alumina moduli usually vary from 2.0 to 3.0 and from 1.0 to 4.0, respectively (Taylor, 1997).
Ali etal. (2013) studied the effect of incorporating up to 2.5% CS in the raw mix by replacing limestone, bauxite and iron ore. An analysis of the LSF and silica and alumina moduli showed that, although the LSF was not affected by the incorporation of CS, the silica and alumina moduli decreased mainly because of the greater Fe2O3 content of CS, when compared to that of the raw ingredients that were replaced. This would suggest that the amount of silicate and aluminate phases would decrease with an increase of the C4AF phase. This was also shown by Bogues method, in which the estimations for the amount of tricalcium silicate (C3S), dicalcium silicate (C2S), tricalcium aluminate (C3A) and tetracalcium aluminoferrite (C4AF) phases were in the ranges of 54.758.9%, 16.319.3%, 4.66.3% and 12.915.3%, respectively, with liquid content varying between 27.4% and 28.9%.
Although the aforementioned methods suggested a decrease of the silicate phases with increasing CS content, the results of the X-ray diffraction analysis indicated that the incorporation of CS resulted in relatively rapid clinker mineral phase formations. Indeed, C3S and C2S contents in samples containing CS, heated at 1400C, were found to be within the range of 5258% and 2328%, respectively, and were comparable to those of the control PC clinker, with C3S and C2S contents of 56% and 26%, respectively, calcined at 1450C.
In the study of Medina etal. (2006), the amount of each phase produced during clinkerisation at 1350 and 1450C was quantified by means of the Rietveld method (Table 5.4). Although no significant changes were found in the C3S phase, a slight increase in the C2S phase was observed when 1.85% CS was used, which would explain the slight decrease in the free CaO content (Table 5.5) in comparison to that of the control PC clinker.
The cement clinker must be correctly burned, to minimise its free lime (CaO) content with the least expenditure of energy (Taylor, 1997). The free lime content of clinker is regarded as a practical measure of the degree of raw mix clinkerisation and is used as a means of controlling the quality of clinker produced. The typical range of free lime content in PC is 0.53%. Table 5.5 presents the free lime content in clinker manufactured with and without CS at different temperatures. As expected, the free lime content decreased with increasing clinkerisation temperatures and it decreased even further when increasing CS was incorporated in the raw mix (Figure 5.4). This trend was explained by the enhanced lime combinability at lower temperatures with the incorporation of CS containing copper oxide (CuO) as well as a decrease in the liquid phases viscosity (Kakali etal., 1996; Kolovos etal., 2005; Ma etal., 2010).
Figure 5.4. Effect of copper slag (CS) incorporation on free lime content of cement clinker subjected to increasing temperatures based on the results of (a) Ali etal. (2013) (b) Supekar (2007) (c) Sahu etal. (2011) (d) Medina etal. (2006) (e) Taeb and Faghihi (2002).
Figure 5.5, which reflects the results presented in Table 5.5, presents the relative free lime of cement clinker samples taken from several studies, in which the CS was ground with the other raw mix components and subjected to normal clinkerisation temperatures (Ali etal., 2013; Medina etal., 2006; Sahu etal., 2011; Supekar, 2007; Taeb and Faghihi, 2002). The results indicate a clear decrease in free CaO content with increasing CS content, revealing greater lime combinability at lower temperatures as observed in other studies (Kakali etal., 1996; Kolovos etal., 2005; Ma etal., 2010). The presence of CuO, which acts both as mineraliser and as flux, decreases the melting temperature by at least 50C and favours the combination of free lime, resulting in accelerated C3S formation (Kolovos etal., 2005).
The phosphate contents of Portland cement clinkers are normally low (around 0.2% as P2O5), although higher levels may be experienced where phosphate is present in significant levels in the raw materials or from alternative fuels such as sewage sludge or meat and bone meal. Fig. 3.2368 shows that up to about 0.5wt% (as P2O5) can be accommodated in the structure of C3S (giving C3S on the diagram) before it decomposes at higher phosphate levels to give a solid solution between C2S and phosphate, and free lime. These products have less satisfactory cementing characteristics. While previous work69 had indicated complete solubility between C2S and C3P at 1500C, later work showed the presence of a miscibility gap, limited by positions PSS and PSS on the figure. The phase diagram shows a number of other solid solutions in addition to the parent phases, C3S, C3P, C4P and C2S. It can be seen that although the primary crystallisation field for C3S extends to 13wt% P2O5, it becomes increasingly narrow towards its maximum limit so that the compositional constraints for the crystallisation of C3S (or C3S) become more stringent. Also, it should be noted that the nearer the bulk composition tends towards the limit of C3S stability, the less C3S can be expected in the clinker. This situation is normally avoided by careful mix proportioning. Consequently, other features such as those appearing in the C4P region of the diagram are less relevant in cement manufacture.
Fig. 3.23. Phase diagram of the system CaOC2SC3P. = phase boundary; = compatibility join at 1500C; PSS, phosphate solid solution at maximum liquidus temperature; PSS, limiting phosphate solid solution at 1500C; PSS, limiting phosphate solid solution at 1500C; C3S, C3S solid solution with Ca2+ and PO43 ions; C4P solid solution with Ca2+ and silicate ions.
The presence of fluoride in phosphatic limestones can have an important influence on the phase equilibria discussed above because coupled substitutions of phosphate and fluoride can occur in C3S. Consideration of the CaOP2O5CaF2 system would show that fluorapatite [3(C3P) CaF2] forms a compatibility with C3S with up to 2mol% of fluorapatite being dissolved in the C3S structure at 1905C. This corresponds to a C3S composition having up to 1.16% P2O5.
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