large breaking ratio gold ore ball mill toggle copper ore on

metallurgist & mineral processing engineer

metallurgist & mineral processing engineer

For its extensive practical experience, 911 Metallurgisthas a clear understanding of what successful mineral processing engineering is and how to go about achieving it. Your goal is the production of a material that is marketable and returns you and your investors sustainable revenues.

Although improvements to the metallurgical processes have been made over the years the fact is that the unit operations, the machines, those too often called black boxes involved have not evolved or changed much since inception. Ore is reduced in size, chemicals are added and minerals separated and upgraded to produce a marketable product. Much of this process is mechanical and generally mistaken for some dark alchemy.We are the Anti-Alchemists.

Our vast experience has been gained through operation and start-up of both small and large scale mining/metallurgical operations in a range of commodities in thebase metals (Cu, Pb, Zn) and theprecious metals (Au, Ag,)

A solid metallurgist understands, the most important aspect of an operating process is its stability. Simple to say, but generally the most ignored in mineral processing. Linked unit operations require each to be stable, and each contains a different set of variables that have to be contended with. Thanks to some degree of stability: operating changes can be made and evaluated; increases in throughput can be made; and equipment performance improved. The more complicated the processes become, the more difficult it is to achieve and maintain stability. In mineral processing, unlike most processing operations, we have limited control of the main input, the feed ore. In most cases this inherently is variable and usually outside of the processors control.

Because you are too close to your own story, you might not see the forest for the trees and have chaos mistaken for stability. We, you, and your group have been battling plant problems for weeks, you start to accept chaos as a daily state of affair and consider it your new stability.

Each mineral processing plant is different: with varied ore types, mining equipment, and management (operating) philosophy. The evaluation and prioritisation of variables that affect the plant performance is the primary function. Implementing changes within the constraints imposed can be difficult, as resources may be limited.

Invariably the ability to solve problems can be confusing due the large numbers of variables that may impact the processes. In most cases problems are not metallurgical in nature but rather operational and mechanical. Problem solving is a process and in many operations this ability is absent. All too often many changes are made together without a solution resulting, on more confusion. Most plants learn to live or survive their problems, not to solve them.

Our engineering team has a global experience in the mining industry across all facets of the mine life-cycle. Our focus is to add value to your project and company by understanding your needs, employing innovative ideas and applying sound engineering while maintaining an economically driven approach. We have a combination of senior level professionals, experienced project managers, and technical staff to execute projects efficiently. We work in a partnership with our clients to achieve their company goals and operational milestones in a timely and cost effective manner.

millstone | create wiki | fandom

millstone | create wiki | fandom

Millstone Renewable Yes Stackable Yes (64) Tools Blast resistance 6 Hardness 1.5 Solid block Yes Full block No Transparent Yes Luminant No Flammable No Catches fire from lava No Rotational Power stats Requires Rotational Power Yes Kinetic Stress impact 4x RPM Speed requirement None

The Millstone is a more basic equivalent of the Crushing Wheels, capable of applying milling recipes to any items inserted from above. A Millstone can be powered by a shaft from below or by a cogwheel from the side.

A Millstone will take some time to process inserted ingredients and will show particles while it is working. Once finished, the Millstone will not drop or output the result by itself, but there are two ways to remove processed items:

ersel ar makine san. ve tic. a

ersel ar makine san. ve tic. a

ERSEL AIR MAKNE SANAY ve TCARET A.. TOSB Otomotiv Yan Sanayi O.S.B. 1.Cadde No:24/2 41420 ayrova / KOCAEL / TURKEY Phone: +90 262 658 13 40 / Fax: +90 262 658 05 27 / E-mail: [email protected]

chest official minecraft wiki

chest official minecraft wiki

A chest can be broken using anything, but an axe is the fastest. Chests always drop themselves when mined. If the chest contains items, the items are also dropped when the chest is broken. If one half of a large chest is destroyed, the corresponding items from the destroyed chest are dropped and the remaining half continues to function as a small chest.

Chests are naturally generated in dungeons, strongholds, jungle temples, desert temples, nether fortresses, villages, end cities, igloos, woodland mansions, shipwrecks, ocean ruins, buried treasure, pillager outposts, bastion remnants, and ruined portals.

The bonus chest is a chest that appears near the player's spawn if the "Bonus chest" option is toggled on the main menu. It generates with a semi-random collection of basic items to help the player survive early on and gather necessary resources, including tools, blocks, and food.

A chest placed adjacent to another chest joins to create a large chest (also known as a double chest). A player can prevent this, and place two small chests side by side, by sneaking while placing the second chest[JE only], pushing a chest into place with a piston[BE only], or placing the second chest facing a different direction from the first chest. Alternatively, trapped chests do not combine with normal chests.

Chests can be moved by pistons,[BE only] and water and lava flow around chests without affecting them. Lava can create fire in air blocks next to chests as if the chests were flammable, but the chests do not actually catch fire and cannot be burned.

A small chest has 27 slots of inventory space, and a large chest has twice that amount, at 54 slots. In the Java Edition interface, the top three rows for a large chest correspond to the right half of the chest when facing it, and the bottom three rows correspond to the left half. In Bedrock Edition, the top three rows correspond to whichever half was placed first and the bottom three to the other half.

Chests open when used. To move items between the chest inventory and the player inventory or hotbar while the chest GUI is open, drag or shift-click the items. Holding Shift and double-clicking while holding an item moves all items of the type clicked on in or out of the chest, to the extent that space is available for them.[Java Edition only] To exit the chest GUI, use the Esc control.

A chest cannot be opened if there is an opaque block above it (in Bedrock Edition, bottom-half slabs prevent chests from being opened). Solid faces do not prevent chests from opening, so the lid of the chest can phase through blocks such as bottom-half slabs, stairs, and transparent full cubes such as glass and ice. Ocelots and cats sitting on chests prevent them from opening. Because chests themselves are functionally transparent, two chests can be stacked on top of one another while still allowing the lower chest to be opened. Players can open chests when players are being hurt (effect of instant damage, poison, wither, fatal poison or on fire) or hurt by anyone (players, zombies, skeletons, etc.).

By default, the GUI of a chest is labeled "Chest" and the GUI of a large chest is labeled "Large Chest". A chest's GUI label can be changed by naming the chest in an anvil before placing it, or by using the /data command[Java Edition only] (for example, to label a chest at (0,64,0) as "Bonus Chest!", use /data merge block 0 64 0 {CustomName:'"Bonus Chest!"'}). If half of a large chest is renamed, that name is used to label the GUI of the entire large chest, but if the named half is destroyed the other half reverts to the default label. If both halves of a large chest have different names, the GUI uses the name of the northernmost or westernmost half.

In Java Edition, a chest can be "locked" by setting its Lock tag using the /data command. If a chest's Lock tag is not blank, the chest cannot be opened unless the player is holding an item with the same name as the Lock tag's text. For example, to lock a chest at (0,64,0) so that the chest cannot be opened unless the player is holding an item named "Chest Key", use /data merge block 0 64 0 {Lock:"Chest Key"}.

The capacity of a chest varies greatly, depending on its size, whether the items inside it are stackable, and whether shulker boxes are used. The minimum capacity is obtained when storing only non-stackable items while the maximum capacity can be achieved when storing items that stack to 64. Filled shulker boxes are not stackable, but each can hold 27 stacks of up to 64 items (excluding other shulker boxes), so filling a chest with them increases the maximum by a factor of 27.

A chest attached to a donkey or mule has only 15 slots. A chest attached to a llama has anywhere from 3 to 15 slots depending upon its "Strength" value (see Llama Data values). The chest cannot be removed except by killing the carrier. The chest can be opened by holding sneak and pressing use, or by riding the carrier and pressing inventory.

If shulker boxes are again used, each donkey, mule or strength value 5 llamas with a chest attached to it can carry up to 405 stacks of items (up to 25920 items), and with strength value 5 llamas, each caravan of 10 llamas with inventories full of shulker boxes can carry up to 4050 stacks of items (up to 259200 items).

This chest can only be seen with the method described on Java Edition, on console, it is in an exclusive texture pack. On December 2426, chests, Ender chests, large chests, and their trapped chest counterparts have their textures changed to "Christmas chests" that look like wrapped Christmas presents. Since the game uses the date shown on the computer, players can access the Christmas chest textures at any time by changing the date and time on their computers to 2426 of December.

rittinger - an overview | sciencedirect topics

rittinger - an overview | sciencedirect topics

As a rule, size reduction operations are heavy in energy consumption (Loncin and Merson, 1979; Hassanpour et al., 2004). As an example, the cost of energy is the single largest item in the total cost of wheat milling. Milling of one bushel (approx. 27kg) of wheat requires 1.74kwh of electric energy (Ryan and Tiffany, 1998).

The total energy consumption of a mill consists of two parts: the energy imparted to the milled material and that needed to overcome friction in bearings and other moving parts of the mill. The energy transferred to the material corresponds to the work of deformation and is stored in the particle as internal stress. When the particle fractures, the stored energy is released. Part of it provides the increment in surface energy resulting from increased surface area but most of it is released as heat. Eventually, friction losses also generate heat. Consequently, size reduction may result in considerable increase in the temperature of the treated material. Temperature rise as a result of size reduction may be an important technological issue, particularly with heat-sensitive products, thermoplastic substances and materials with high fat content. When necessary, this problem is addressed by air- or water-cooling of the machine or using cryogenics such as liquid nitrogen (cryo-milling).

Mechanical efficiency m of a size reduction device is defined as the ratio of the energy transferred to the material to the total energy consumption W of the device, per unit mass of material treated.

A different expression for energy requirement of size reduction has been proposed by Kick. Kick assumes that the energy needed to reduce the size of the material by a certain proportion (say by half or by one order of magnitude) is constant (first order relationship). Kick's law is written as follows:

Sugar crystals were ground from an average Sauter diameter of 500m to powder with an average Sauter diameter of 100m. The net energy consumption was 0.5kWh per ton. What would be the net energy consumption for grinding the crystals to 50m powder:a.according to Rittinger's lawb.according to Kick's law.

a.Rittinger's law:E=K(1x2-1x1)K is calculated from the first milling data and applied to the second milling:K = 0.5/(1/1001/500) = 62.5kwh.m/tonE = 62.5 (1/501/500) = 1.125kwh/tonb.Kick's law:E=Klog(x1/x2)K = 0.5/log (100/500) = 0.715E = 0.715* log (50/500) = 0.715kwh/ton

Bond [BON 52] made the hypothesis that the exponent of dp is the arithmetic mean between Rittinger and Kicks values, being 0.5. In addition, it returns energy not to the volume but to the mass being processed and, for the specific energy, he writes:

Bond calls Wi the work index, literally meaning energy index. Naturally, Wi depends only, in principle, on the nature of the milled body. It is the reason why he provides Wi values (as well as for the true density) for most commonly used ores in Table IIIA on page 548 of [BON 61b].

In Table IIIA, Bond [BON 61b, p. 548] provides the standard specific energies expressed in kilowatt hour which he calls a short ton (whose value is 2,000 pounds).kWhshortton=3600kilojoule0.90718metircton=3.968.103kilojoule.ton11kWhshortton1=3.968kJ.kg1

The concept of measuring Wi is the following. We put a total mass MT of ore to be milled in a small laboratory grinder. After a determined number of rotations NT, we empty the grinder and sort the mass MT on a sieve with an opening P1. The mass of the underflow is MP=MTMR, where MR is the mass of overflow. We complete the mass MR with fresh ore to obtain MT that we reload into the grinder.

Assuming that the mass MT placed in the grinder has, at the start of each of its last cycles, 80% underflow with a size of df microns and that this same mass has, at the end of each of these cycles, 80% underflow with a size of dp microns, Bonds standard specific energy Wi will be:

Then, 700g of this output, compacted following a standard procedure, are dry milled in the vessel (=305mm and L=305mm). The rotation speed of the vessel is 70 rev.mn1, indicating 85% of the critical speed. The balls load weighs 20.125kg and is made up of a specific number of balls with sizes ranging from 12.7 to 38mm.

The method consists of grinding the load for a short period of time (100300 revolutions). We then sieve the load with a screen size xT set in advance (that is, 300m for example) and replace the downflow with an equal mass of fresh feed. We repeat the operation until there is a constant ratio equal to 2.5 between the overflow and the underflow and the mass MF of underflow obtained for the content of the mill is equal to a constant.

The length of time required for each operation must be determined by trial and error and the number of operations may, depending on the situation, vary from 7 to 15. The method is therefore not simple.

For a ball mill, if the reduction ratio becomes less than 3 (target grinding of concentrates), the energy index Wi must be multiplied by a given coefficient given by the authors equation 27 [BON 61b, p. 545].

If we want to reduce a very large compact solid (df=) into particles with diameter dp=100m, the required specific energy becomes equal to Bonds energy index Wi that is measured in kilowatt hour per short ton (907.18kg) of solid. Values for this energy index will be provided for many ores [BON 60].

Note that carbon is missing from Bonds list. This issue was dealt with by Chandler [CHA 65]. The applicable standard method for carbon is Hardgroves [HAD 32] test that Chandler [CHA 65] describes.Note IIn practice, it is very difficult to make use of a bulk solid when the size of all its particles is equal to dp. This is why Bond defines the sizes df and dp as screen openings that let through 80% of the feeds solid mass and grinder output, respectively.Note IIBond [BON 54] proposed correspondent relationships between his energy index and magnitudes characterizing the capacity during grinding given in other texts written by Bond himself [BON 49].

In practice, it is very difficult to make use of a bulk solid when the size of all its particles is equal to dp. This is why Bond defines the sizes df and dp as screen openings that let through 80% of the feeds solid mass and grinder output, respectively.

This coefficient is only applicable when R<6.ExampleLet us consider by dry grinding 40 ton.h1 of ore in a ball mill that enters at 5mm and exits at 200m. The apparent density of the ore is 1.5 and its energy index Ei is equal to 15. We estimate that the diameter of the mill is less than 3.8m.C1=1.3C2=1C3=2.4430.2=0.96xo=40001315=3724mC4=1+15750003724372450.2=1.11C5=1C7=1E=151.110020010050001.30.961.11E=22.860.710.14=13kWh/tonPa=4013=520kW

Let us consider by dry grinding 40 ton.h1 of ore in a ball mill that enters at 5mm and exits at 200m. The apparent density of the ore is 1.5 and its energy index Ei is equal to 15. We estimate that the diameter of the mill is less than 3.8m.C1=1.3C2=1C3=2.4430.2=0.96xo=40001315=3724mC4=1+15750003724372450.2=1.11C5=1C7=1E=151.110020010050001.30.961.11E=22.860.710.14=13kWh/tonPa=4013=520kW

We could use the same reasoning for a rod mill but the energy needed to raise a rod will be proportional to DCB2LD and the number of rods will be DCB2LD and the number of rods would be D2L/DCB2L . We obtain:

There is a need to describe the relationship between the capacity of the mill and the properties of the milled material. Appropriate methods are based on various comminution theories, the most common of which are Rittingers [2], Kicks [3] and Bonds [4].

The commonly used method to evaluate the grindability of coal in medium speed pulverizers is the Hardgrove Grindability Index (HGI) [5]. The HGI test is based on Rittingers theory. It allows to predict the mill output, performance and energy requirements, and (qualitatively) also the particle size distribution after milling [6]. As the value of HGI increases, the capacity of the mill increases as well. Numerous experiences show that if the HGI test is a good indicator of milling performance for medium speed mills when grinding coal, it is poor for other materials such as biomass. Another disadvantage of HGI is that the tester is a batch device and does not reflect the continuous grinding process.

Broad dissemination of biomass burning in PF boilers caused the search for other indicators better reflecting the comminution of such materials [79]. The studies show that in this case better results give the methods based on Bonds theory.

Although it is impossible to estimate accurately the amount of energy required in order to effect a size reduction of a given material, a number of empirical laws have been proposed. The two earliest laws are due to Kick(7) and von Rittinger(8), and a third law due to Bond(9,10) has also been proposed. These three laws may all be derived from the basic differential equation:

which is known as Kick's law. This supposes that the energy required is directly related to the reduction ratio L1/L2 which means that the energy required to crush a given amount of material from a 50 mm to a 25 mm size is the same as that required to reduce the size from 12 mm to 6 mm. In equations 2.3 and 2.4, KR and KK are known respectively as Rittinger's constant and Kick's constant. It may be noted that neither of these constants is dimensionless.

Neither of these two laws permits an accurate calculation of the energy requirements. Rittinger's law is applicable mainly to that part of the process where new surface is being created and holds most accurately for fine grinding where the increase in surface per unit mass of material is large. Kick's law, more closely relates to the energy required to effect elastic deformation before fracture occurs, and is more accurate than Rittinger's law for coarse crushing where the amount of surface produced is considerably less.

Bond terms Ei the work index, and expresses it as the amount of energy required to reduce unit mass of material from an infinite particle size to a size L2 of 100 m, that is q = . The size of material is taken as the size of the square hole through which 80 per cent of the material will pass. Expressions for the work index are given in the original papers(8,9) for various types of materials and various forms of size reduction equipment.

Hardgrove Indexbased on Rittinger's Law, which states that the power consumption is proportional to the new surface created. A prepared sample receives a definite amount of grinding energy in laboratory ring-roll pulverizer. The sample is compared with a coal chosen as having 100 grindability (Pittsburgh Seam coal).Index=136.93w where w = wt of material passing 200 B.S. sieve (obtained from orig. wt of 50 g-wt retained on sieve). Usual range of indices 25 to 75. For details see A.S.T.M. D409.

Finally, a look should be taken at coal elasticity, hardness and strength. However, a particular matter of importance which arises from those consideration is the ease of coal grinding, an important step in whatever coal preparation efforts for further processing. The more fundamental material properties are covered reasonably by Berkowitz (1994), so the discussion here will be limited to coal grindability. For that purpose, use is made of two different indices, both determined experimentally with the material to be ground. One is the Hardgrove grindability index and the other the Bond work index.

The Hardgrove index is determined using the ASTM method D 40971. It involves grinding 50g of the material, e.g. coal, of specified size (1630 mesh cut) in a specified ball-and-race mill for 60 revolutions. The amount of 200 mesh material is measured (w grams) and the index is defined as I = 13+ 6.93w. Thus, the higher the index, the easier is the grinding task. This method loosely assumes that the specific energy consumed is proportional to the new surface generated, following the concept of Rittingers law of comminution.

Berkowitz (1994 p.96) gives a generalized variation of the Hardgrove index with coal rank. According to the variation, anthracites are hard to grind, bituminous coals the easiest, and the subbituminous more difficult, with lignites down to the same low index level as anthracites. It is suggested that the decrease in the index below daf coal of 85% is caused by plastic deformation and aggregation of the softer coal particles, hence reducing the 200 mesh fraction generated by the grinding test.

The Bond work index (Bond, 1960) is based on Bonds law, which states that the energy consumed is proportional to the 1.5 power of particle size rather than the square of Rittingers law. Accordingly, the energy consumed in reducing the particle size from xF to xp (both measured as 80% undersize) is given by

We should note that the higher the value of the work index, the more difficult it is to grind the material. A compilation of data is available, for example, in Perrys Chemical Engineers Handbook (Perry et al., 1984). For coal, one average value is given, with Ei = 11.37 for = 1.63. Bonds law is useful because of the extensive comparative database.

Interestingly, Hukki (1961) offers a Solomonic settlement between the different grinding theories (rather than laws). A great deal of additional material related to grinding, or size reduction, comminution, is available in handbooks, e.g. by Prasher (1987) and research publications in journals such as Powder Technology. A very brief overview of grinding equipment is given in Section 1.5.3.

This family of models is the oldest of the comminution models and they continue to find widespread use (Morrell, 2014a). Energy-based models assume a relationship between energy input of the comminution device and the resultant effective particle size of the product. Many rely on the feed and product size distributions being self-similar; that is, parallel when cumulative finer is plotted in log-log space (Chapter 4). The energy input is for net power, that is, after correcting for motor efficiency and drive train mechanical losses. Typically, energy is measured as kWh t1 or Joules, depending on the model.

The oldest theory, Von Rittinger (1867), stated that the energy consumed in size reduction is proportional to the area of new surface produced. The surface area of a known weight of particles of uniform diameter is inversely proportional to the diameter, hence Von Rittingers law equates to:

As Lynch and Rowland (2005) note, the means to make measurements of energy and size necessary to validate the Von Rittinger and Kick models did not exist until the middle of the twentieth century when electrical motors and precision laboratory instruments became available. The literature from this period includes work by a group at the Allis Chalmers Company who were trying to calibrate Von Rittingers equation to industrial rod mills (Bond and Maxson, 1938; Myers et al., 1947).

Often referred to as the third theory, Bond (1952) stated that the energy input is proportional to the new crack tip length produced in particle breakage. Bond redefined his theory to rather be an empirical relationship in a near-final treatise (Bond, 1985). The equation is commonly written as:

where W is the energy input (work) in kilowatt hours per metric ton (or per short ton in Bonds original publications), Wi is the work index (or Bond work index) in kilowatt hours per metric ton, and P80 and F80 are the 80% product and feed passing sizes, in micrometers.

Solving Eq. (5.1b) for n=3/2 gives the same form as Eq. (5.4) with the constant 2 K ahead of the bracket. In effect the 2 K is replaced by (10Wi), which is convenient because Wi becomes equal to W in the case of grinding from a theoretical infinite feed size to 80% passing 100m. The Bond model remains the most widely used, at least for the conventional comminution equipment in use at the time Bond developed the model and calibrated it against industrial data. It is one reason that the 80% passing size became the common single point metric (mean) of a particle size distribution.

A modification of Eq. (5.1a,b) was proposed by Hukki (1962), namely substituting n by a function of particle size, f(x). This provoked debate over the size range that the three established models applied to. What can be agreed is that all the models predict that energy consumption will increase as product particle size (i.e., P) decreases. Typical specific energy values (in kWh t1) are (Morrell, 2014b): primary crushing (i.e., 1000-100mm), 0.1-0.15; secondary crushing (100-10mm), 1-1.2; coarse grinding (10-1mm), 3-3.5; and fine grinding (1-0.1mm), 10.

Fine grinding tests are sometimes expressed as a signature plot (He et al., 2010), which is an experimentally fitted version of Eq. (5.1a,b) with n=f(x). A laboratory test using a fine grinding mill is conducted where the energy consumption is carefully measured and a slurry sample is extracted periodically to determine the 80% passing size. The energy-time relationship versus size is then plotted and fitted to give (in terms of Eq. (5.1a,b)) a coefficient K and a value for the exponent f (x).

The problem that occurs when trying to solve Eq. (5.5) is the variable nature of the function g(x). A pragmatic approach was to assume M is a constant over the normal range of particle sizes treated in the comminution device and leave the variation in size-by-size hardness to be taken up by f(x). Morrell (2009) gives the following:

where Mi is the work index parameter related to the breakage property of an ore and the type of comminution machine, W is the specific comminution energy (kWh t1), P and F are the product and feed 80% passing size (m), and f(x) is given by (Morrell, 2006):

The parameter Mi takes on different values depending on the comminution machine: Mia for primary tumbling mills (AG/SAG mills) that applies above 750m; Mib for secondary tumbling mills (e.g., ball mills) that applies below 750m; Mic for conventional crushers; and Mih for HPGRs. The values for Mia, Mic, and Mih were developed using the SMC Test combined with a database of operating comminution circuits. A variation of the Bond laboratory ball work index test was used to determine values of Mib. This is similar to the approach Bond used in relating laboratory results to full scale machines. The methodology continues to be refined as the database expands (Morrell, 2010).

Morrell (2009) gave a worked example comparing the energy requirements for three candidate circuits to illustrate the calculations. Taking just the example for the fine particle tumbling mill serves that purpose here (Example 5.1).

From the Mi data the relevant value is Mib=18.8kWh t1. Noting it is fine grinding then the feed F80 is taken as 750m. Combining Eqs. (5.6) and (5.7) and substituting the values:W=18.84(106(0.295+750/1,000,000)750(0.295+750/1,000,000))=8.4(kWht1)

Milling or grinding can reduce coarse fly ash particles to fine particle size in a similar range as particles separated out from air classification. But milling does not directly compete with air classification, as milling is not a filtration technique. Milling reduces the fly ash particle size by breaking up large spherical particles into smaller irregularly shaped particles that can have a negative impact on rheology. Milling also consumes more energy than air classification to obtain the fine particle size distribution and has a size limitation, as it is difficult to reduce the particle sizes down to less than 10m. The advantage of milling is that the entire milled quantity will consist of only one defined particle size distribution instead of fine and coarse particle size distributions.

Milling breaks solid materials into smaller pieces by grinding, crushing, or cutting by attrition, collision, or compressive forces. For particles less than 50m, the energy needed to grind the material down to the desired size follows the Von Rittinger grinding law, while the particle sizes less than 50mm but greater than 50m follows the Bond grinding law, as listed here:

Three types of millings have been used to grind fly ash to smaller particle size: ball milling, vibration milling, and plate (pan or plane) milling. Ball milling is typically loaded with particulate materials at its 30%40% capacity. A higher rotation speed, longer processing time, greater ball density, or greater impact force produces finer particle size distribution. The accumulated volumes of Class II fly ash, which has no more than 25% by weight higher than 45m, and its ball-milled samples at different times of 15, 30, 45, 60, 90, and 120min are shown in Fig. 10.12. A longer grinding time produces finer particle size. After 15min of grinding, all milled fly ashes become Class I fly ash, which is no more than 12% by weight greater than 45m, according to fly ash classification under the Chinese standard, GB/T 15962005.

Vibration milling is the core technology for energetically modified cement (EMC) technology, which was patented in 1993 by Dr. Vladimir Ronin. Plate milling was introduced in 2015 at the World of Coal Ash conference by Professor Li Hui from Xi`an University of Architecture and Technology. Her paper showed the energy consumption to grind down fly ash with D50 of 21.57m down to 3m is 1019, 1323, and 120kWh/ton for ball, vibration, and plate milling, respectively.

Air-dried limestone was crushed continuously in a laboratory single-toggle Blake jaw crusher designed to provide a throw of 228mm. The lower opening close set was 102mm and the maximum bottom opening was 330mm. The gape was 813mm and the width of hopper 1067mm. 90% of the ore commenced crushing 200mm from the bottom of the crusher. The Bond index was estimated as 15 kWh/t. Assuming that the density of the limestone was 2.6t/m3, determine1.the optimum RPM of the toggle,2.the maximum annual capacity of crusher with 99% availability,3.power consumption at the optimum speed.

A single toggle Blake jaw crusher with 22.8cm 47.7cm receiving hopper crushed gold ore at the rate of 85t/h with closed setting at 2.54cm and maximum opening of 3.8cm. The work index of the ore was 13.5kWh/t.

The feed to a jaw crusher was 60mm+40mm and the product analysed:Screen Size (mm)Product (% Retained)Screen Size (mm)Product (% Retained)810.0+0.3510.1+421.8+0.255.5+216.3+0.1256.2+0.7520.10.12510.0

The compressive strength of the mineral was 20MN/m2. The crusher was next used to crush a second mineral of compressive strength 55MN/m2 at 5kg/s. The feed size of the second mineral was 55+40mm and yielded a product whose average size was 0.4mm. Estimate the change in power required during the second operation.

A Blake jaw crusher had the following dimensions: Gape=160cm, open set=24.4cm, close set=5.0cm. The width of the hopper was 1.5 times the gape. The ore contained 20% material minus 4.0cm. The bulk density of the rock was 1.75t/m3 and the nip angle 22.8.

A cement manufacturer needed to produce lime at the rate of 140,000t/year in a rotary kiln operating 360days in the year. Limestone for the purpose contained 30% CaO. The S.G. of limestone was 2.7. The mined material had a top size of 40 100cm after screening through a grizzly. The kiln accepted top size of 10cm. A single toggle Blake jaw crusher was available for crushing. Assume the shape factor of the feed and the product were the same.

A jaw crusher was used to crush a chert ore. The top size of the ore was 25cm and the moisture content was less than 3%. It was required to produce a product 100% of which would be less than 4cm. The shape factor of feed and product was 1.7. Assume that the cumulative weight-size curve was a straight line, determine:1.crusher size,2.rate of crushing (QT).

A jaw crusher had a gape of 685mm. It was charged continuously by a conveyor belt to keep a charge level constant at 46cm from the bottom of the jaws. A reduction ratio of 7.5 was desired. If the maximum opening between the jaws at the discharge end was fixed at 20cm for a material of density 2.8, compute1.the angle between the crusher faces (assume flat),2.operating speed and critical speed of operation,3.the rate of crushing when the angle between plates is increased by 2.

The angle between the straight faces of a Blake jaw crusher was progressively altered from 22 to 28 in steps of 2. 1200kg of a material of bulk density 1460kg/m3 was crushed each time. Indicate1.the adjustments to the set that would be necessary each time to maintain same production rate,2.the mathematical relation between the angle of nip and the set.

Iron ore was crushed in a jaw crusher. The average sizes of the feed (F80) and product (P80) were 50 and 10mm, respectively. The energy consumed during crushing was found to be 5kWh/t. The top size of the material was then altered to an average size (F80) of 75mm when the product size (P80) of 5mm was required. Estimate the energy to crush in the altered condition.

A single toggle jaw crusher crushed limestone having an average size of 75mm. The size analysis of the product wasSize (mm)Mass % RetainedSize (mm)Mass % Retained12.50.201.515.07.58.00.755.25.051.00.402.12.513.00.205.5

The closed set of an operating jaw crusher was 125mm. A continuous stream of ore was fed at the rate of 30t/h. On an average, 10% of the ore was less than the set. The F80 was 410mm in size. The crusher was initially operated at 200rpm at a reduction ratio of 1:4, but the toggle speed was to be increased. Calculate1.the maximum speed, C, at which it can be operated,2.the maximum capacity at the maximum operating speed of the toggle.

tnt official minecraft wiki

tnt official minecraft wiki

Once activated, TNT turns into an entity, which includes being affected by gravity. The new primed TNT is spawned at the center (+0.5,+0.5,+0.5) of where the TNT block was, like a cube with an edge length of 0.98. Its fuse lasts 40 redstone ticks (4 seconds/80 game ticks) if activated by redstone or fire, or a random number between 10 and 30 game ticks (0.5 to 1.5 seconds) if it is destroyed by another explosion.

Once spawned, primed TNT is given a vertical velocity of 0.2blocks per tick, and a horizontal velocity of 0.02blocks per tick in a random direction. Given these velocities, the TNT travels 0.166 blocks (or 6 block pixels) horizontally before it stops, given there is no block in the way. When the countdown timer expires, the TNT explodes. If in the air, TNT falls roughly 77 blocks before exploding once it is ignited. The explosion has an explosive force of 4.

Primed TNT's texture blinks, alternating every 0.5 seconds between the TNT block's texture, and a copy of it that has been brightened to near-white. The effect is dynamic and the brightened texture can't be found in the assets.

When primed TNT detonates while in water or lava, it does not break any blocks. It does still damage players, mobs, and other entities.[Java Edition only][1] Primed TNT that detonates outside water can still damage submerged blocks.

To make TNT destroy blocks in the water, e.g. to enter an ocean monument from the top, one can place sand or gravel on the TNT before it is primed. Priming the TNT causes the sand or gravel to fall one block, engulfing the TNT. Because the TNT is no longer immersed in water, it can destroy the surrounding blocks.

Primed TNT is not teleported when entering a nether portal; instead, it passes through portal blocks.[Java Edition only] Primed TNT teleports as expected when entering an end portal, maintaining its direction and speed. The fuse timer keeps counting down (unless the spawn chunks are unloaded, then it pauses until a player loads the chunks).

If the TNT is primed atop any sort of fence post that is two blocks high or larger, it falls through the fence block on which it was activated and stops on the next lower one. Its detonation damages only the fence block it was "stuck" in.

TNT can be used in a variety of traps. The simplest of them a land mine is made of TNT connected to a pressure plate or tripwire with redstone. One such trap generates naturally in desert pyramids under the loot area.

gyratory crusher - an overview | sciencedirect topics

gyratory crusher - an overview | sciencedirect topics

Gyratory crushers were invented by Charles Brown in 1877 and developed by Gates around 1881 and were referred to as a Gates crusher [1]. The smaller form is described as a cone crusher. The larger crushers are normally known as primary crushers as they are designed to receive run-on-mine (ROM) rocks directly from the mines. The gyratory crushers crush to reduce the size by a maximum of about one-tenth its size. Usually, metallurgical operations require greater size reduction; hence, the products from the primary crushers are conveyed to secondary or cone crushers where further reduction in size takes place. Here, the maximum reduction ratio is about 8:1. In some cases, installation of a tertiary crusher is required where the maximum reduction is about 10:1. The secondary crushers are also designed on the principle of gyratory crushing, but the construction details vary.

Similar to jaw crushers, the mechanism of size reduction in gyratory crushers is primarily by the compressive action of two pieces of steel against the rock. As the distance between the two plates decreases continuous size reduction takes place. Gyratory crushers tolerate a variety of shapes of feed particles, including slabby rock, which are not readily accepted in jaw crushers because of the shape of the feed opening.

The gyratory crusher shown in Figure 2.6 employs a crushing head, in the form of a truncated cone, mounted on a shaft, the upper end of which is held in a flexible bearing, whilst the lower end is driven eccentrically so as to describe a circle. The crushing action takes place round the whole of the cone and, since the maximum movement is at the bottom, the characteristics of the machine are similar to those of the Stag crusher. As the crusher is continuous in action, the fluctuations in the stresses are smaller than in jaw crushers and the power consumption is lower. This unit has a large capacity per unit area of grinding surface, particularly if it is used to produce a small size reduction. It does not, however, take such a large size of feed as a jaw crusher, although it gives a rather finer and more uniform product. Because the capital cost is high, the crusher is suitable only where large quantities of material are to be handled.

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%.

Maintenance of the wear components in both gyratory and cone crushers is one of the major operating costs. Wear monitoring is possible using a Faro Arm (Figure 6.10), which is a portable coordinate measurement machine. Ultrasonic profiling is also used. A more advanced system using a laser scanner tool to profile the mantle and concave produces a 3D image of the crushing chamber (Erikson, 2014). Some of the benefits of the liner profiling systems include: improved prediction of mantle and concave liner replacement; identifying asymmetric and high wear areas; measurement of open and closed side settings; and quantifying wear life with competing liner alloys.

Crushers are widely used as a primary stage to produce the particulate product finer than about 50100mm. They are classified as jaw, gyratory, and cone crushers based on compression, cutter mill based on shear, and hammer crusher based on impact.

A jaw crusher consists essentially of two crushing plates, inclined to each other forming a horizontal opening by their lower borders. Material is crushed between a fixed and a movable plate by reciprocating pressure until the crushed product becomes small enough to pass through the gap between the crushing plates. Jaw crushers find a wide application for brittle materials. For example, they are used for comminution of porous copper cake. A Fritsch jaw crusher with maximal feed size 95mm, final fineness (depends on gap setting) 0.315mm, and maximal continuous throughput 250Kg/h is shown in Fig. 2.8.

A gyratory crusher includes a solid cone set on a revolving shaft and placed within a hollow body, which has conical or vertical sloping sides. Material is crushed when the crushing surfaces approach each other and the crushed products fall through the discharging opening.

Hammer crushers are used either as a one-step primary crusher or as a secondary crusher for products from a primary crusher. They are widely used for crushing hard metal scrap for different hard metal recycling processes. Pivoted hammers are pendulous, mounted on the horizontal axes symmetrically located along the perimeter of a rotor. Crushing takes place by the impact of material pieces with the high speed moving hammers and by contact with breaker plates. A cylindrical grating or screen is placed beneath the rotor. Materials are reduced to a size small enough to pass through the openings of the grating or screen. The size of the product can be regulated by changing the spacing of the grate bars or the opening of the screen.

The feature of the hammer crushers is the appearance of elevated pressure of air in the discharging unit of the crusher and underpressure in the zone around the shaft close to the inside surface of the body side walls. Thus, the hammer crushers also act as high-pressure, forced-draught fans. This may lead to environmental pollution and product losses in fine powder fractions. A design for a hammer crusher (Fig. 2.9) essentially allows a decrease of the elevated pressure of air in the crusher discharging unit [5]. The A-zone beneath the screen is communicated through the hollow ribs and openings in the body side walls with the B-zone around the shaft close to the inside surface of body side walls. As a result, the circulation of suspended matter in the gas between A and B zones is established and the high pressure of air in the discharging unit of crusher is reduced.

Crushers are widely used as a primary stage to produce the particulate product finer than about 50100 mm in size. They are classified as jaw, gyratory and cone crushers based on compression, cutter mill based on shear and hammer crusher based on impact.

A jaw crusher consists essentially of two crushing plates, inclined to each other forming a horizontal opening by their lower borders. Material is crushed between a fixed and a movable plate by reciprocating pressure until the crushed product becomes small enough to pass through the gap between the crushing plates. Jaw crushers find a wide application for brittle materials. For example, they are used for comminution of porous copper cake.

A gyratory crusher includes a solid cone set on a revolving shaft and placed within a hollow body, which has conical or vertical sloping sides. Material is crushed when the crushing surfaces approach each other and the crushed products fall through the discharging opening.

Hammer crushers are used either as a one-step primary crusher or as a secondary crusher for products from a primary crusher. They are widely used for crushing of hard metal scrap for different hard metal recycling processes.

Pivoted hammers are pendulous, mounted on the horizontal axes symmetrically located along the perimeter of a rotor and crushing takes place by the impact of material pieces with the high speed moving hammers and by contact with breaker plates. A cylindrical grating or screen is placed beneath the rotor. Materials are reduced to a size small enough pass through the openings of the grating or screen. The size of product can be regulated by changing the spacing of the grate bars or the opening of the screen.

The feature of the hammer crushers is the appearance of elevated pressure of air in the discharging unit of the crusher and underpressure in the zone around of the shaft close to the inside surface of the body side walls. Thus, the hammer crushers also act as high-pressure forced-draught fans. This may lead to environmental pollution and product losses in fine powder fractions.

A design for a hammer crusher (Figure 2.6) allows essentially a decrease of the elevated pressure of air in the crusher discharging unit [5]. The A-zone beneath the screen is communicated through the hollow ribs and openings in the body side walls with the B-zone around the shaft close to the inside surface of body side walls. As a result, circulation of suspended matter in the gas between A- and B-zones is established and high pressure of air in the discharging unit of crusher is reduced.

Jaw crushers are mainly used as primary crushers to produce material that can be transported by belt conveyors to the next crushing stages. The crushing process takes place between a fixed jaw and a moving jaw. The moving jaw dies are mounted on a pitman that has a reciprocating motion. The jaw dies must be replaced regularly due to wear. Figure 8.1 shows two basic types of jaw crushers: single toggle and double toggle. In the single toggle jaw crusher, an eccentric shaft is installed on the top of the crusher. Shaft rotation causes, along with the toggle plate, a compressive action of the moving jaw. A double toggle crusher has, basically, two shafts and two toggle plates. The first shaft is a pivoting shaft on the top of the crusher, while the other is an eccentric shaft that drives both toggle plates. The moving jaw has a pure reciprocating motion toward the fixed jaw. The crushing force is doubled compared to single toggle crushers and it can crush very hard ores. The jaw crusher is reliable and robust and therefore quite popular in primary crushing plants. The capacity of jaw crushers is limited, so they are typically used for small or medium projects up to approximately 1600t/h. Vibrating screens are often placed ahead of the jaw crushers to remove undersize material, or scalp the feed, and thereby increase the capacity of the primary crushing operation.

Both cone and gyratory crushers, as shown in Figure 8.2, have an oscillating shaft. The material is crushed in a crushing cavity, between an external fixed element (bowl liner) and an internal moving element (mantle) mounted on the oscillating shaft assembly. An eccentric shaft rotated by a gear and pinion produces the oscillating movement of the main shaft. The eccentricity causes the cone head to oscillate between the open side setting (o.s.s.) and closed side setting (c.s.s.). In addition to c.s.s., eccentricity is one of the major factors that determine the capacity of gyratory and cone crushers. The fragmentation of the material results from the continuous compression that takes place between the mantle and bowl liners. An additional crushing effect occurs between the compressed particles, resulting in less wear of the liners. This is also called interparticle crushing. The gyratory crushers are equipped with a hydraulic setting adjustment system, which adjusts c.s.s. and thus affects product size distribution. Depending on cone type, the c.s.s. setting can be adjusted in two ways. The first way is by rotating the bowl against the threads so that the vertical position of the outer wear part (concave) is changed. One advantage of this adjustment type is that the liners wear more evenly. Another principle of setting adjustment is by lifting/lowering the main shaft. An advantage of this is that adjustment can be done continuously under load. To optimize operating costs and improve the product shape, as a rule of thumb, it is recommended that cones always be choke-fed, meaning that the cavity should be as full of rock material as possible. This can be easily achieved by using a stockpile or a silo to regulate the inevitable fluctuation of feed material flow. Level monitoring devices that detect the maximum and minimum levels of the material are used to start and stop the feed of material to the crusher as needed.

Primary gyratory crushers are used in the primary crushing stage. Compared to the cone type crusher, a gyratory crusher has a crushing chamber designed to accept feed material of a relatively large size in relation to the mantle diameter. The primary gyratory crusher offers high capacity thanks to its generously dimensioned circular discharge opening (which provides a much larger area than that of the jaw crusher) and the continuous operation principle (while the reciprocating motion of the jaw crusher produces a batch crushing action). The gyratory crusher has capacities starting from 1200 to above 5000t/h. To have a feed opening corresponding to that of a jaw crusher, the primary gyratory crusher must be much taller and heavier. Therefore, primary gyratories require quite a massive foundation.

The cone crusher is a modified gyratory crusher. The essential difference is that the shorter spindle of the cone crusher is not suspended, as in the gyratory, but is supported in a curved, universal bearing below the gyratory head or cone (Figure 8.2). Power is transmitted from the source to the countershaft to a V-belt or direct drive. The countershaft has a bevel pinion pressed and keyed to it and drives the gear on the eccentric assembly. The eccentric assembly has a tapered, offset bore and provides the means whereby the head and main shaft follow an eccentric path during each cycle of rotation. Cone crushers are used for intermediate and fine crushing after primary crushing. The key factor for the performance of a cone type secondary crusher is the profile of the crushing chamber or cavity. Therefore, there is normally a range of standard cavities available for each crusher, to allow selection of the appropriate cavity for the feed material in question.

Depending on the size of the debris, it may either be ready to enter the recycling process or need to be broken down to obtain a product with workable particle sizes, in which case hydraulic breakers mounted on tracked or wheeled excavators are used. In either case, manual sorting of large pieces of steel, wood, plastics and paper may be required, to minimise the degree of contamination of the final product.

The three types of crushers most commonly used for crushing CDW materials are the jaw crusher, the impact crusher and the gyratory crusher (Figure 4.4). A jaw crusher consists of two plates, with one oscillating back and forth against the other at a fixed angle (Figure 4.4(a)) and it is the most widely used in primary crushing stages (Behera etal., 2014). The jaw crusher can withstand large and hard-to-break pieces of reinforced concrete, which would probably cause the other crushing machines to break down. Therefore, the material is initially reduced in jaw crushers before going through any other crushing operation. The particle size reduction depends on the maximum and minimum size of the gap at the plates (Hansen, 2004).

An impact crusher breaks the CDW materials by striking them with a high-speed rotating impact, which imparts a shearing force on the debris (Figure 4.4(b)). Upon reaching the rotor, the debris is caught by steel teeth or hard blades attached to the rotor. These hurl the materials against the breaker plate, smashing them into smaller particle sizes. Impact crushers provide better grain-size distribution of RA for road construction purposes, and they are less sensitive to material that cannot be crushed, such as steel reinforcement.

Generally, jaw and impact crushers exhibit a large reduction factor, defined as the ratio of the particle size of the input to that of the output material. A jaw crusher crushes only a small proportion of the original aggregate particles but an impact crusher crushes mortar and aggregate particles alike and thus generates a higher amount of fine material (OMahony, 1990).

Gyratory crushers work on the same principle as cone crushers (Figure 4.4(c)). These have a gyratory motion driven by an eccentric wheel. These machines will not accept materials with a large particle size and therefore only jaw or impact crushers should be considered as primary crushers. Gyratory and cone crushers are likely to become jammed by fragments that are too large or too heavy. It is recommended that wood and steel be removed as much as possible before dumping CDW into these crushers. Gyratory and cone crushers have advantages such as relatively low energy consumption, a reasonable amount of control over the particle size of the material and production of low amounts of fine particles (Hansen, 2004).

For better control of the aggregate particle size distribution, it is recommended that the CDW should be processed in at least two crushing stages. First, the demolition methodologies used on-site should be able to reduce individual pieces of debris to a size that the primary crusher in the recycling plant can take. This size depends on the opening feed of the primary crusher, which is normally bigger for large stationary plants than for mobile plants. Therefore, the recycling of CDW materials requires careful planning and communication between all parties involved.

A large proportion of the product from the primary crusher can result in small granules with a particle size distribution that may not satisfy the requirements laid down by the customer after having gone through the other crushing stages. Therefore, it should be possible to adjust the opening feed size of the primary crusher, implying that the secondary crusher should have a relatively large capacity. This will allow maximisation of coarse RA production (e.g., the feed size of the primary crusher should be set to reduce material to the largest size that will fit the secondary crusher).

The choice of using multiple crushing stages mainly depends on the desired quality of the final product and the ratio of the amounts of coarse and fine fractions (Yanagi etal., 1998; Nagataki and Iida, 2001; Nagataki etal., 2004; Dosho etal., 1998; Gokce etal., 2011). When recycling concrete, a greater number of crushing processes produces a more spherical material with lower adhered mortar content (Pedro etal., 2015), thus providing a superior quality of material to work with (Lotfi etal., 2017). However, the use of several crushing stages has some negative consequences as well; in addition to costing more, the final product may contain a greater proportion of finer fractions, which may not always be a suitable material.

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. [4]). 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 [5].

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.

Beneficiation of copper ores is done almost exclusively by selective froth flotation. Flotation entails first attaching fine copper mineral particles to bubbles rising through an orewater pulp and, second, collecting the copper minerals at the top of the pulp as a briefly stable mineralwaterair froth. Noncopper minerals do not attach to the rising bubbles; they are discarded as tailings. The selectivity of the process is controlled by chemical reagents added to the pulp. The process is continuous and it is done on a large scale103 to 105 tonnes of ore feed per day.

Beneficiation is begun with crushing and wet-grinding the ore to typically 10100m. This ensures that the copper mineral grains are for the most part liberated from the worthless minerals. This comminution is carried out with gyratory crushers and rotary grinding mills. The grinding is usually done with hard ore pieces or hard steel balls, sometimes both. The product of crushing and grinding is a waterparticle pulp, comprising 35% solids.

Flotation is done immediately after grindingin fact, some flotation reagents are added to the grinding mills to ensure good mixing and a lengthy conditioning period. The flotation is done in large (10100m3) cells whose principal functions are to provide: clouds of air bubbles to which the copper minerals of the pulp attach; a means of overflowing the resulting bubblecopper mineral froth; and a means of underflowing the unfloated material into the next cell or to the waste tailings area.

Selective attachment of the copper minerals to the rising air bubbles is obtained by coating the particles with a monolayer of collector molecules. These molecules usually have a sulfur atom at one end and a hydrophobic hydrocarbon tail at the other (e.g., potassium amyl xanthate). Other important reagents are: (i) frothers (usually long-chain alcohols) which give a strong but temporary froth; and (ii) depressants (e.g., CaO, NaCN), which prevent noncopper minerals from floating.

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