Hello, my partner! Let's explore the mining machine together!

[email protected]

# coal crushing plant purchase 1 hour 20

## jaw crusher - an overview | sciencedirect topics

The mechanism of movement of rocks down the crusher chamber determines the capacity of jaw crushers. The movement can be visualised as a succession of wedges (jaw angles) that reduce the size of particles progressively by compression until the smaller particles pass through the crusher in a continuous procession. The capacity of a jaw crusher per unit time will therefore depend on the time taken for a particle to be crushed and dropped through each successive wedge until they are discharged through the bottom. The frequency of opening and closing of the jaws, therefore, exerts a significant action on capacity.

Following the above concepts, several workers, such as Hersam [6]. Gaudin [7], Taggart [8], Rose and English [9], Lynch [3], Broman [10], have attempted to establish mathematical models determining the capacity.

Although it is not truly applicable to hard rocks, for soft rocks it is reasonably acceptable [1]. This expression, therefore, is of limited use. The expressions derived by others are more appropriate and therefore are discussed and summarised here.

Rose and English [9] determined the capacity of a jaw crusher by considering the time taken and the distance travelled by the particles between the two plates after being subjected to repeat crushing forces between the jaws. Therefore, dry particles wedged between level A and level B (Figure4.4) would leave the crusher at the next reverse movement of the jaw. The maximum size of particle dropping out of the crusher (dMAX) will be determined by the maximum distance set at the bottom between the two plates (LMAX). The rate at which the crushed particles pass between the jaws would depend on the frequency of reversal of the moving jaw.

The distance, h, between A and B is equal to the distance the particle would fall during half a cycle of the crusher eccentric, provided the cycle frequency allows sufficient time for the particle to do so. If is the number of cycles per minute, then the time for one complete cycle is [60/] seconds and the time for half a cycle is [60/2]. Thus, h, the greatest distance through which the fragments would fall freely during this period, will be

Then for a fragmented particle to fall a distance h in the crusher, the frequency must be less than that given by Equation (4.10). The distance h can be expressed in terms of LMIN and LMAX, provided the angle between the jaws, , is known. From Figure4.4, it can be seen that

Rose and English [9] observed that with increasing frequency of the toggle movement the production increased up to a certain value but decreased with a further increase in frequency. During comparatively slower jaw movements and frequency, Rose and English derived the capacity, QS, as

Equation (4.12) indicates that the capacity, QS, is directly proportional to frequency. At faster movement of the jaws where the particle cannot fall the complete distance, h, during the half cycle, QF was found to be inversely proportional to frequency and could be expressed by the relation

The relationship between the frequency of operation and capacity of the jaw crusher can be seen in Figure4.5. This figure is plotted for values of LT=0.228m, W=1.2m, LMIN=0.10m, R=10, G=1 and the value of varied between 50 and 300rpm.

It should be noted that while considering the volume rates, no consideration was made to the change of bulk density of the material or the fractional voidage. However, during the crushing operation the bulk density of the ore changes as it passes down the crusher. The extent of the change depends on

PK is considered a size distribution function and is related to capacity by some function (PK). As the particles decrease in size, while being repeatedly crushed between the jaws, the amount of material discharged for a given set increases. Rose and English related this to the set opening and the mean size of the particles that were discharged. Defining this relation as it can be written as

The capacity is then dependant on some function which may be written as (). Equations (4.16) and (4.17) must, therefore, be incorporated into the capacity equation. Expressing capacity as mass of crusher product produced per unit time, capacity can be written as

The bulk density of the packing will depend on the particle size distribution. The relation between PK and (PK) and and () is shown in Figure4.6. It is based on a maximum possible bulk density of 40%.

As the closed set size must be less than the feed size, () may be taken as equal to 1 for all practical purposes. The maximum capacity of production can be theoretically achieved at the critical speed of oscillation of the moving jaw. The method of determining the critical speed and maximum capacity is described in Section4.2.3

The capacity of a jaw crusher is given by the amount of crushed material passing the discharge opening per unit time. This is dependent on the area of the discharge opening, the properties of the rock, moisture, crusher throw, speed, nip angle, method of feeding and the amount of size reduction.

In order to calculate the capacity of crushers, Taggart [8] considered the size reduction, R80, as the reduction ratio of the 80% passing size of the feed, F80, and product, P80. This may be written as

Hersam [6] showed that at a fixed set and throw, a decrease in feed size reduced the reduction ratio and increased the tonnage capacity. A fraction of the crusher feed is usually smaller than the minimum crusher opening at the discharge end (undersize) and, therefore, passes through the crusher without any size reduction. Thus, as the feed size decreases, the amount actually crushed becomes significantly less than the total feed. The crusher feed rate can increase to maintain the same crushing rate. Taggart expressed the relationship between crusher capacity and reduction ratio in terms of a reduction ton or tonne, QR defined as

The reduction tonnage term is dependent on the properties of the material crushed so that for a given reduction ratio, the crusher capacity will vary for different materials. Taggart attempted to compensate for this by introducing the comparative reduction tonne, QRC, which is related to the reduction tonne by the expression

The comparative reduction tonne is a standard for comparison and applies for the crushing conditions of uniform full capacity feeding of dry thick bedded medium-hard limestone where K=1. The factor K is determined for different conditions and is a function of the material crushability (kC), moisture content (kM) and crusher feeding conditions (kF). K is expressed as

To evaluate K, the relative crushability factor, kC, of common rocks was considered and is given in Table4.2. In the table, the crushability of limestone is considered standard and taken as equal to 1.

The moisture factor, kM, has little effect on primary crushing capacities in jaw crushers and could be neglected. However when clay is present or the moisture content is high (up to 6%) sticking of fine ores on the operating faces of the jaws is promoted and will reduce the production rate. The moisture effect is more marked during secondary crushing, where a higher proportion of fines are present in the feed.

The feed factor kF, applies to the manner in which the crusher is fed, for example, manually fed intermittently or continuously by a conveyor belt system. In the latter case, the rate of feeding is more uniform. The following values for factor kF are generally accepted:

The reduction ratio of the operation is estimated from screen analysis of the feed and product. Where a screen analysis is not available, a rough estimate can be obtained if the relation between the cumulative mass percent passing (or retained) for different size fractions is assumed to be linear (Figure4.7).

Figure4.7 is a linear plot of the scalped and unscalped ores. The superimposed data points of a crusher product indicate the fair assumption of a linear representation. In the figure, a is the cumulative size distribution of the unscalped feed ore (assumed linear) and b is the cumulative size distribution of the scalped ore. xS is the aperture of the scalping screen and d1 and d2 are the corresponding sizes of the scalped and unscalped feed at x cumulative mass percentage. Taking x equal to 20% (as we are required to estimate 80% that is passing through), it can be seen by simple geometry that the ratio of the 80% passing size of the scalped feed to the 80% passing size of the unscalped feed is given by

Run of mine granite is passed through a grizzly (45.7cm) prior to crushing. The ore is to be broken down in a jaw crusher to pass through a 11.5cm screen. The undersize is scalped before feeding to the jaw crusher. Assuming the maximum feed rate is maintained at 30t/h and the shapes of feed and product are the same and the crusher set is 10cm, estimate the size of jaw crusher required and the production rate.

Substituting values, assuming cubic-shaped particles where the shape factor=1.7, we haveF80=0.81.745.7+0.210=64.15cmandP80=0.81.711.5=15.64cmR80=64.1515.64=4.10HenceQRC=22.744.100.64=145.4t/h

For a jaw crusher the thickness of the largest particle should not normally exceed 8085% of the gape. Assuming in this case the largest particle to be crushed is 85% of the gape, then the gape of the crusher should be=45.7/0.85=53.6cm and for a shape factor of 1.7, the width should be=45.7 1.7=78cm.

From the data given by Taggart (Figure4.8), a crusher of gape 53.6cm would have a comparative reduction tonnage of 436 t/h. The corresponding crushing capacity would beQT=4360.644.10=68.1t/hand is thus capable of handling the desired capacity of 22.74 t/h.

To determine the capacity of jaw and gyratory crushers, Broman [10] divided the crusher chamber into different sections and determined the volume of each section in terms of the angle that the moving jaw subtended with the vertical. Broman suggested that the capacity per stroke crushed in each section would be a function of the top surface and the height of the section. Referring to Figure4.9, if is the angle of nip between the crusher jaws and LT and LMAX are the throw and open side setting, respectively, then

Michaelson [8] expressed the jaw crusher capacity in terms of the gravity flow of a theoretical ribbon of rock through the open set of the crusher times a constant, k. For a rock of SG 2.65, Michaelsons equation is given as

For a set of crusher sizes and set openings, the calculations obtained from the work of Rose and English and others can be compared with data from equipment manufacturers. Figure4.10 shows a plot of the results. Assuming a value of SC of 1.0, the calculations show an overestimation of the capacity recommended by the manufacturers. As Rose and English pointed out, the calculation of throughput is very dependent on the value of SC for the ore being crushed. The diagram also indicates that the calculations drop to within the installed plant data for values of SC below 1.0. Most other calculation methods tend to estimate higher throughputs than the manufacturers recommend; hence, the crusher manufacturers should always be consulted.

The Values Used in the Calculation were 2.6 SG, (PK)=0.65, ()=1.0 and SC=0.51.0 (R&E); k=0.4 (Hersam); k=0.3 (Michaelson); k=1.5 (Broman) and =275rpm. The Max and Min Lines Represent the Crushers Nominal Operating Capacity Range.

Jaw crushers are heavy-duty machines and hence must be robustly constructed. The main frame is often made from cast iron or steel, connected with tie-bolts. It is commonly made in sections so that it can be transported underground for installation. Modern jaw crushers may have a main frame of welded mild steel plate.

The jaws are usually constructed from cast steel and fitted with replaceable liners, made from manganese steel, or Ni-hard, a Ni-Cr alloyed cast iron. Apart from reducing wear, hard liners are essential to minimize crushing energy consumption by reducing the deformation of the surface at each contact point. The jaw plates are bolted in sections for simple removal or periodic reversal to equalize wear. Cheek plates are fitted to the sides of the crushing chamber to protect the main frame from wear. These are also made from hard alloy steel and have similar lives to the jaw plates. The jaw plates may be smooth, but are often corrugated, the latter being preferred for hard, abrasive ores. Patterns on the working surface of the crushing members also influence capacity, especially at small settings. The corrugated profile is claimed to perform compound crushing by compression, tension, and shearing. Conventional smooth crushing plates tend to perform crushing by compression only, though irregular particles under compression loading might still break in tension. Since rocks are around 10 times weaker in tension than compression, power consumption and wear costs should be lower with corrugated profiles. Regardless, some type of pattern is desirable for the jaw plate surface in a jaw crusher, partly to reduce the risk of undesired large flakes easily slipping through the straight opening, and partly to reduce the contact surface when crushing flaky blocks. In several installations, a slight wave shape has proved successful. The angle between the jaws is usually less than 26, as the use of a larger angle causes particle to slip (i.e., not be nipped), which reduces capacity and increases wear.

In order to overcome problems of choking near the discharge of the crusher, which is possible if fines are present in the feed, curved plates are sometimes used. The lower end of the swing jaw is concave, whereas the opposite lower half of the fixed jaw is convex. This allows a more gradual reduction in size as the material nears the exit, minimizing the chance of packing. Less wear is also reported on the jaw plates, since the material is distributed over a larger area.

The speed of jaw crushers varies inversely with the size, and usually lies in the range of 100350rpm. The main criterion in determining the optimum speed is that particles must be given sufficient time to move down the crusher throat into a new position before being nipped again.

The throw (maximum amplitude of swing of the jaw) is determined by the type of material being crushed and is usually adjusted by changing the eccentric. It varies from 1 to 7cm depending on the machine size, and is highest for tough, plastic material and lowest for hard, brittle ore. The greater the throw the less danger of choking, as material is removed more quickly. This is offset by the fact that a large throw tends to produce more fines, which inhibits arrested crushing. Large throws also impart higher working stresses to the machine.

In all crushers, provision must be made for avoiding damage that could result from uncrushable material entering the chamber. Many jaw crushers are protected from such tramp material (often metal objects) by a weak line of rivets on one of the toggle plates, although automatic trip-out devices are now common. Certain designs incorporate automatic overload protection based on hydraulic cylinders between the fixed jaw and the frame. In the event of excessive pressure caused by an overload, the jaw is allowed to open, normal gap conditions being reasserted after clearance of the blockage. This allows a full crusher to be started under load (Anon., 1981). The use of guard magnets to remove tramp metal ahead of the crusher is also common (Chapters 2 and 13Chapter 2Chapter 13).

Jaw crushers are supplied in sizes up to 1,600mm (gape)1,900mm (width). For coarse crushing application (closed set~300mm), capacities range up to ca. 1,200th1. However, Lewis et al. (1976) estimated that the economic advantage of using a jaw crusher over a gyratory diminishes at crushing rates above 545th1, and above 725th1 jaw crushers cannot compete.

In hardening and martempering conditions austenitic manganese steel was free from carbides both at the grain boundaries and in the grains. Hence, the crusher jaws produced with austenitic manganese in these conditions eradicated brittle failure experienced in locally produced crusher jaws.

Hardening followed by tempering precipitated carbide at the grain boundaries and in the grains instead of reducing the residual stress associated with hardening. The volume fraction of these carbides, however, increased with tempering temperature.

In martempering conditions austenitic manganese steel had better plastic flows due to a decrease in overall thermal gradient and reduction in residual stresses associated with heat-treatment operations. This gave a better combination of hardness and toughness than austenitic manganese steel in hardening conditions used for the production of imported crusher jaws.

Srikanth [7] used a jaw crusher to create37m coal dust particles. Coal samples were obtained from coal mines in addition to some samples from the same source as Thakur's samples. They used a Microtrac Standard Range Analyzer (SRA) and Small Particle Analyser (SPA), which measured projected area (and hence diameter) using laser scattering and diffraction, respectively. The data were combined and plotted on a RosinRammler graph (discussed in Chapter 8). Their main findings were as follows:

Higher rank coals produced more total dust (<15m) and respirable dust (<7m). Semianthracite coal produced 3.7 times more total dust and 4.2 times more respirable dust compared with high-volatile bituminous coal.

The RosinRammler graph distribution parameter, n, was also rank dependent. The value for n was 0.68, 0.84, 0.90, and 0.95 for semianthracite, low-volatile coal, high-volatile bituminous coal, and subbituminous coals, respectively. This is similar to findings by Thakur (refer to Chapter 8 in the book).

A material is crushed in a Blake jaw crusher such that the average size of particle is reduced from 50 mm to 10 mm with the consumption of energy of 13.0 kW/(kg/s). What would be the consumption of energy needed to crush the same material of average size 75 mm to an average size of 25 mm:

The size range involved by be considered as that for coarse crushing and, because Kick's law more closely relates the energy required to effect elastic deformation before fracture occurs, this would be taken as given the more reliable result.

In an investigation by the U.S. Bureau of Mines(14), in which a drop weight type of crusher was used, it was found that the increase in surface was directly proportional to the input of energy and that the rate of application of the load was an important factor.

This conclusion was substantiated in a more recent investigation of the power consumption in a size reduction process which is reported in three papers by Kwong et al.(15), Adams et al.(16) and Johnson etal.(17). A sample of material was crushed by placing it in a cavity in a steel mortar, placing a steel plunger over the sample and dropping a steel ball of known weight on the plunger over the sample from a measured height. Any bouncing of the ball was prevented by three soft aluminium cushion wires under the mortar, and these wires were calibrated so that the energy absorbed by the system could be determined from their deformation. Losses in the plunger and ball were assumed to be proportional to the energy absorbed by the wires, and the energy actually used for size reduction was then obtained as the difference between the energy of the ball on striking the plunger and the energy absorbed. Surfaces were measured by a water or air permeability method or by gas adsorption. The latter method gave a value approximately double that obtained from the former indicating that, in these experiments, the internal surface was approximately the same as the external surface. The experimental results showed that, provided the new surface did not exceed about 40 m2/kg, the new surface produced was directly proportional to the energy input. For a given energy input the new surface produced was independent of:

Between 30 and 50 per cent of the energy of the ball on impact was absorbed by the material, although no indication was obtained of how this was utilised. An extension of the range of the experiments, in which up to 120 m2 of new surface was produced per kilogram of material, showed that the linear relationship between energy and new surface no longer held rigidly. In further tests in which the crushing was effected slowly, using a hydraulic press, it was found, however, that the linear relationship still held for the larger increases in surface.

In order to determine the efficiency of the surface production process, tests were carried out with sodium chloride and it was found that 90 J was required to produce 1 m2 of new surface. As the theoretical value of the surface energy of sodium chloride is only 0.08 J/m2, the efficiency of the process is about 0.1 per cent. Zeleny and Piret(18) have reported calorimetric studies on the crushing of glass and quartz. It was found that a fairly constant energy was required of 77 J/m2 of new surface created, compared with a surface-energy value of less than 5 J/m2. In some cases over 50 per cent of the energy supplied was used to produce plastic deformation of the steel crusher surfaces.

The apparent efficiency of the size reduction operation depends on the type of equipment used. Thus, for instance, a ball mill is rather less efficient than a drop weight type of crusher because of the ineffective collisions that take place in the ball mill.

Further work(5) on the crushing of quartz showed that more surface was created per unit of energy with single particles than with a collection of particles. This appears to be attributable to the fact that the crushing strength of apparently identical particles may vary by a factor as large as 20, and it is necessary to provide a sufficient energy concentration to crush the strongest particle. Some recent developments, including research and mathematical modelling, are described by Prasher(6).

The main sources of RA are either from construction and ready mixed concrete sites, demolition sites or from roads. The demolition sites produce a heterogeneous material, whereas ready mixed concrete or prefabricated concrete plants produce a more homogeneous material. RAs are mainly produced in fixed crushing plant around big cities where CDWs are available. However, for roads and to reduce transportation cost, mobile crushing installations are used.

The materiel for RA manufacturing does not differ from that of producing NA in quarries. However, it should be more robust to resist wear, and it handles large blocks of up to 1m. The main difference is that RAs need the elimination of contaminants such as wood, joint sealants, plastics, and steel which should be removed with blast of air for light materials and electro-magnets for steel. The materials are first separated from other undesired materials then treated by washing and air to take out contamination. The quality and grading of aggregates depend on the choice of the crusher type.

Jaw crusher: The material is crushed between a fixed jaw and a mobile jaw. The feed is subjected to repeated pressure as it passes downwards and is progressively reduced in size until it is small enough to pass out of the crushing chamber. This crusher produces less fines but the aggregates have a more elongated form.

Hammer (impact) crusher: The feed is fragmented by kinetic energy introduced by a rotating mass (the rotor) which projects the material against a fixed surface causing it to shatter causing further particle size reduction. This crusher produces more rounded shape.

However, the gyratory crusher is sensitive to jamming if it is fed with a sticky or moist product loaded with fines. This inconvenience is less sensitive with a single-effect jaw crusher because mutual sliding of grinding surfaces promotes the release of a product that adheres to surfaces.

The profile of active surfaces could be curved and studied as a function of the product in a way to allow for work performed at a constant volume and, as a result, a higher reduction ratio that could reach 20. Inversely, at a given reduction ratio, effective streamlining could increase the capacity by 30%.

The theoretical work of Rose and English [11] to determine the capacity of jaw crushers is also applicable to gyratory crushers. According to Rose and English, Equation (5.4) can be used to determine the capacity, Q, of gyratory crushers:

Capacities of gyratory crushers of different sizes and operation variables are published by various manufacturers. The suppliers have their own specifications which should be consulted. As a typical example, gyratory crusher capacities of some crushers are shown in Tables5.5 and 5.6.

About 100g heavy metal contaminated construction and demolition (C&D) waste is weighed and preliminarily crushed by a jaw crusher. Then the crushed C&D waste is mixed well and reduced by quartering twice. After that, the sample is dried at 100C for 1h. An electromagnetic crusher is used as a fine crushing for about 46min. Crushed sample is placed in a polypropylene screw-cap plastic bottles for storage.

Teflon crucibles used for digestion should be soaked in 1:1 nitric acid for 12h, wash with distilled water, and dry for later use. Volumetric flasks should be soaked in 1:1 nitric acid for 12h and washed with distilled water.

Before digestion, 0.10000.3000g of C&D waste powder is accurately weighed and evenly spread on the bottom of Teflon crucibles. Then they are placed in oven and dried for 2h at 120C together till constant weight. Aqua regia (18mL) (hydrochloric acid:nitric acid=3:1) is added, and 2mL 40% hydrofluoric acid is added 10min later. The crucibles with lids on are placed on an electric heating plate at 180C and heated till the solid waste is dissolved. Then, 30mL deionized water is added and the heating should be continuously maintained till the solution is vaporized to 23mL. Transfer the liquid to a 25mL plastic volumetric flask after it is cooled down, in which the volumetric flask should be washed with 1% nitric acid solution three times. Add deionized water to a certain volume and filter through 0.22m membrane. Place the solution at 4C for analysis.