## how to use an enema for clearing the bowel

Robert Burakoff, MD, MPH, is board-certified in gastroentrology. He is the vice chair for ambulatory services for the department of medicine at Weill Cornell Medical College in New York, where he is also a professor. He was the founding editor and co-editor in chief of Inflammatory Bowel Diseases.

An enema is the introduction of liquid, most often mineral oil, through the anus and into the large intestine. An enema may be given to treat constipation, to administer medication or barium, or as part of the procedure to empty the contents of the bowel before a test (such as acolonoscopy prep).

Using homemade enemas are not recommended, nor is using an enema containing liquids or substances other than what is recommended by a physician. There is no evidence that using an enema for "detoxing" or for reasons other than cleaning the bowel before a test or procedure or for removing impacted stool has any health benefits.

An enema that one buys in the pharmacy has a nozzle on the end of a small bag. The bag is filled with the liquid that's to be injected into the body. The nozzle is inserted into the anus and the bag is squeezed, sending the liquid out of the nozzle and into the last part of the colon (the rectum).

The liquid is usually held in the rectum for some specified amount of time. It could just be held until the urge to move the bowels comes on. In some cases, it might be suggested that the enema is held inside the body for a few minutes or longer.

In some cases, the liquid used in an enema is just salt water, and in other, it contains a laxative. Check with your doctor if you are unsure as to which type of enema is recommended for you. Here are some common types of enema liquids:

In the treatment of some types of conditions, including inflammatory bowel disease (IBD), medication might be given with an enema. Rowasa (a 5-aminosalicylic drug), used to treat ulcerative colitis, is given this way.

This is usually to treat inflammation that is found in the last section of the colon, where the enema liquid will reach, but that might not be the case in all uses. It's usually recommended that these enemas are used at night, and that instead of releasing the bowels, that the enema is held in all night, to give the medication time to work.

There are practitioners who offer enemas with other substances in them (coffee, lemon juice, milk), which are claimed to offer some health benefits. Use of these types of enemas without the supervision of a physician is not approved or recommended to treat any condition.

People with IBD should be especially wary of these types of enemas. There is the potential to introduce harmful materials into the body with the use of a nonmedical enema. In addition, the colon contains various forms of beneficial bacteria, and the use of an enema may disrupt the bacterial flora and cause harm.

Enemas are not harmlessand should be used only on the advice of a physician. Using enemas on a regular basis can have an effect on the muscles in the colon. In time, the muscles will stop working properly to move stool along, which worsens the problems with constipation.

Richter JM, Arshi NK, Oster G. Oral 5-Aminosalicylate, Mesalamine Suppository, and Mesalamine Enema as Initial Therapy for Ulcerative Proctitis in Clinical Practice with Quality of Care Implications.Can J Gastroenterol Hepatol. 2016;2016:6928710. doi:10.1155/2016/6928710

## bond work index - an overview | sciencedirect topics

The Bond work index is not solely a material constant but is influenced by the grinding conditions. For example, the finer the grind size desired, the higher is the kWh/t required to grind to that size. Magdalinovic [38] measured the Bond work index of three ore types using different test screen sizes. He produced a correlation between the mass of test screen undersize per revolution, G, and the square root of the test screen size, D:

The constant K2 is also dependent on ore type and ranged from 1.4 to 1.5. A regression of Magdalinovics data including the feed 80% passing size gives an average value of 1.485 for K2. If we extend this relationship to any sample of screened material then this gives an approximate estimate of the 80% passing size as 67.3% of the top size. This compares with a value of 66.7% of the 99% passing size obtained from data in Table3.3.

Using Magdalinovics method, from the results of a Bond work index test at a single test screen size, the constants K1 and K2 can be calculated and from these values, the work index at any test screen size can be estimated.

An alternative approach to determine the effect of closing screen size on the Bond ball mill work index (BWi), in the absence of extensive test work, is to use computer simulation. The batch grinding process has been modelled using the sizemass balance approach (Austin [37], Chapter11) and if we can do this, then we can effectively simulate the Bond ball mill work index test. Yan and Eaton [39] measured the selection function and breakage distribution parameters for the Austin grinding model and demonstrated the BWi simulation with soft and medium/hard ore samples. The measured BWi was 14.0 and 6.6kWh/t for the medium/hard and soft ore, respectively, at a closing screen size of 106 m compared with the simulated values of 13.2 and 5.6kWh/t.

The ability to simulate the Bond work index test also allows examination of truncated ball mill feed size distributions on the work index. For grinding circuits where the feed to a ballmill is sent directly to the classifier and the cyclone underflow feeds the ball mill (see Figure3.10), a question arises as to whether this practice will alter the ball mill work index (BWi) of the material being ground and hence have an impact on the energy used in the mill for grinding. Some might conclude that a higher percentage of coarse material in the mill feed will increase the amount of material that needs to be ground to produce the end product and hence it will affect the BWi. Others, in the absence of contrary evidence, assume that there is no change in the work index. Figure3.11 shows the typical circuit represented by the standard Bond work index correlation and Figure3.10 represents the scalped or truncated feed case.

The procedure for the work index test bases the BWi value on the calculation of new fines generated in the test. This means that the fraction of fines in the feed should not influence the test result significantly, if at all. For example, for a sample with 20% of 300 m material in the feed, if this is not scalped out of the fresh feed, then the mill charge, at 250% circulating load will contain 0.2/3.5 or 5.7% of 300 m in the mill charge compared with 0% for a scalped fresh feed, at a closing screen of 300 m. This should not have a great influence on the production of new fines unless the test was carried out in a wet environment and the fines contained a high percentage of clays to affect the viscosity of the grind environment. Thus for a Bond test (dry test), the difference between the scalped and unscalped BWi result is expected to be minor. In a plant operation where the environment is wet and clays are present, a different result may be observed.

Tests carried out to confirm this have clouded the water a little. Three rock types were tested with scalped and unscalped feeds with two samples showing higher BWi values for the scalped ore and the other sample showing a lower value [40].

In the work index test simulation, it is easy to change the closing screen size to examine the effect on the BWi. The results of such a simulation are shown in Figure3.12 where the simulated test was performed at different closing screen sizes and different scalping sizes. This shows that for scalping sizes at or below the closing screen size of the test, the BWi values are not affected. The scalping size of zero refers to the un-scalped mill feed. For scalped screen sizes above the closing screen size, the BWi values start to increase. The increase in BWi is more pronounced at the larger closing screen sizes. At a closing screen size of 300 m and a scalped size of 600 m, the increase in BWi is 4%.

Another outcome of the simulation is the effect of the closing screen size on the work index. As the closing size decreases, the ore must be ground finer, using more energy and producing a higher work index. Further simulations at even larger closing screen sizes show the BWi to increase. This dip in BWi with closing screen size has been observed experimentally, as shown in Figure3.13, with the minimum in BWi occurring at different closing screen sizes for different rock types [41,42].

Bond impact crushability work index (CWi) (Bond, 1963) results reported for iron ores vary from hard iron ore (17.7kWh/t) to medium hardness iron ore (11.3kWh/t) and friable iron ore (6.3kWh/t) (Table 2.11; Clout et al., 2007). The CWi for hard iron ores typically overlaps with those reported for BIF (taconite) iron ores while the range in values in Table 2.11 covers that for different types of iron ores and materials reported earlier by Bond (1963), with some relevant data in Table 2.12.

The most widely used parameter to measure ore hardness is the Bond work index Wi. Calculations involving Bonds work index are generally divided into steps with a different Wi determination for each size class. The low energy crushing work index laboratory test is conducted on ore specimens larger than 50mm, determining the crushing work index (WiC, CWi or IWi (impact work index)). The rod mill work index laboratory test is conducted by grinding an ore sample prepared to 80% passing 12.7mm ( inch, the original test being developed in imperial units) to a product size of approximately 1mm (in the original and still the standard, 14 mesh; see Chapter 4 for definition of mesh), thus determining the rod mill work index (WiR or RWi). The ball mill work index laboratory test is conducted by grinding an ore sample prepared to 100% passing 3.36mm (6 mesh) to product size in the range of 45-150m (325-100 mesh), thus determining the ball mill work index (WiB or BWi). The work index calculations across a narrow size range are conducted using the appropriate laboratory work index determination for the material size of interest, or by chaining individual work index calculations using multiple laboratory work index determinations across a wide range of particle size.

To give a sense of the magnitude, Table 5.1 lists Bond work indices for a selection of materials. For preliminary design purposes such reference data are of some guide but measured values are required at the more advanced design stage.

A major use of the Bond model is to select the size of tumbling mill for a given duty. (An example calculation is given in Chapter 7.) A variety of correction factors (EF) have been developed to adapt the Bond formula to situations not included in the original calibration set and to account for relative efficiency differences in certain comminution machines (Rowland, 1988). Most relevant are the EF4 factor for coarse feed and the EF5 factor for fine grinding that attempt to compensate for sizes ranges beyond the bulk of the original calibration data set (Bond, 1985).

The standard Bond tumbling mill tests are time-consuming, requiring locked-cycle testing. Smith and Lee (1968) used batch-type tests to arrive at the work index; however, the grindability of highly heterogeneous ores cannot be well reproduced by batch testing.

Berry and Bruce (1966) developed a comparative method of determining the hardness of an ore. The method requires the use of a reference ore of known work index. The reference ore is ground for a certain time (T) in a laboratory tumbling mill and an identical weight of the test ore is then ground for the same time. Since the power input to the mill is constant (P), the energy input (E=PT) is the same for both reference and test ore. If r is the reference ore and t the ore under test, then we can write from Bonds Eq. (5.4):

Work indices have been obtained from grindability tests on different sizes of several types of equipment, using identical feed materials (Lowrison, 1974). The values of work indices obtained are indications of the efficiencies of the machines. Thus, the equipment having the highest indices, and hence the largest energy consumers, are found to be jaw and gyratory crushers and tumbling mills; intermediate consumers are impact crushers and vibration mills, and roll crushers are the smallest consumers. The smallest consumers of energy are those machines that apply a steady, continuous, compressive stress on the material.

A class of comminution equipment that does not conform to the assumption that the particle size distributions of a feed and product stream are self-similar includes autogenous mills (AG), semi-autogenous (SAG) mills and high pressure grinding rolls (HPGR). Modeling these machines with energy-based methods requires either recalibrating equations (in the case of the Bond series) or developing entirely new tests that are not confused by the non-standard particle size distributions.

Variability samples must be tested for the relevant metallurgical parameters. Ball mill design requires a Bond work index, BWi, for ball mills at the correct passing size; SAG mill design requires an appropriate SAG test, for example, SPI (Chapter 5). Flotation design needs a valid measure of kinetics for each sample, including the maximum attainable recovery and rate constants for each mineral (Chapter 12). Take care to avoid unnecessary testing for inappropriate parameters, saving the available funds for more variability samples rather than more tests on few samples. Remember that it must be possible to use the measured values for the samples to estimate the metallurgical parameters for the mine blocks in order to describe the ore body, and these estimates will be used in process models to forecast results for the plant. Always include some basic mineralogical examination of each sample.

The expression for computing the power consumption (P) derived theoretically by Rose and English [9] involved the knowledge of Bonds work index (Wi). To evaluate the work index they considered the maximum size in the feed and also the maximum size of particles in the discharge from the crusher. To determine the size through which 80% of the feed passed, they considered a large database relating the maximum particle size and the undersize. From the relation it was concluded that F80 was approximately equal to 0.7 times the largest size of particle. Taking the largest size of the particle that should be charged to a jaw crusher as 0.9 times the gape, F80 was written as

Also, to establish the P80 from the largest product size, Rose and English considered that the largest particle size discharged from the bottom of the crusher would occur at the maximum open set position and hence

For operating a jaw crusher it is necessary to know the maximum power required consistently with the reduction ratio and the gape and closed side settings. The maximum power drawn in a system will occur at the critical speed. Thus for maximum power, Q in Equation (4.51) is replaced with QM from Equation (4.19) to give

The largest size of ore pieces mined measured 560mm (average) and the smallest sizes averaged 160mm. The density of the ore was 2.8t/m3. The ore had to be crushed in a C-63 type jaw crusher 630 440. At a reduction ratio of 4, 18% of the ore was below the maximum size required. Determine:1.the maximum operating capacity of the crusher,2.the optimum speed at which it should be operated.

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.

Rock fragmentation is a consequence of unstable extension of multiple cracks. Theoretically, rock fragmentation is also a facture mechanics problem. Two major differences between rock fracture and rock fragmentation are that (1) rock fragmentation deals with many cracks, but rock fracture deals with only one or a few, and (2) rock fragmentation concerns the size distribution of the fragments produced, but rock fracture does not. There are two important factors in rock fragmentation: (1) total energy consumed and (2) size distribution of fragments. In a study on crushing and grinding, fracture toughness has been taken as a key index similar to the Bond Work Index. Due to many cracks dealt with, rock fragmentation is a very complicated and difficult fracture problem. To achieve a good fragmentation, we need to know how the energy is distributed, which factors influence energy distribution, what is the size distribution, and so on. In practice such as mining and quarrying, it is of importance to predict and examine size distribution so as to make fragmentation optimized by modifying the blast plan or changing the fragmentation system. About size distribution, there are a number of distribution functions such as Weibulls distribution function [11], Cunninghams Kuz-Ram model [12], and the Swebrec function [13]. In engineering practice, how to develop a feasible and simple method to judge rock fragmentation in the field is still a challenging but significant job and will be in the future.

Although the fracture toughness of a rock is very important in rock fracture, the strengths of the rock are also useful in rock engineering. In the following we will see that the strengths and fracture toughness of a rock have a certain relation with each other, partly because of a similar mechanism in the micro-scale failure.

Bong's Work Index is used in Bong's law of comminution energy. It states that the total work useful in breakage is inversely proportional to the length of the formed crack tips and directly proportional to the square root of the formed surface:

where W is the specific energy expenditure in kilowatt-hours per ton and dp and df are the particle size in microns at which 80% of the corresponding product and feed passes through the sieve; CB is a constant depending on the characteristics of materials; and Sp and Sf are the specific surface areas of product and initial feed, respectively. Wi is called Bond's Work Index in kilowatt-hours per ton. It is given by the empirical equation:

where P1 is the sieve opening in microns for the grindability test, Gb.p. (g/rev) is the ball mill grindability, dp is the product particle size in microns (80% of product finer than size P1 passes) and df is the initial feed size in microns (80% of feed passes). A standard ball mill is 305mm in internal diameter and 305mm in internal length charged with 285 balls, as tabulated in Table 2.1. The lowest limit of the total mass of balls is 19.5Kg. The mill is rotated at 70 rev/min. The process is continued until the net mass of undersize produced by revolution becomes a constant Gb.p in the above equation.

To investigate the influence of the coal type on the stampability factor K, stamping tests with eight different coals (C1C8 in Table11.1) were carried out, using the Hardgrove grindability index (HGI) as a measure for the material dependency. The grindability is broadly defined as the response of a material to grinding effort. It can be interpreted as the resistance of the material against particularization. It is not an absolutely measurable physical property of the material. Generally, grindability can be determined either based on product constant fineness method (Bond work index Wi) or on constant useful grinding work method (HGI). The correlation between HGI and Wi can be described by the formula (11.5):

HGI is influenced by the petrographic composition of coal. HGI was developed to find a relationship between petrographic properties and strength of coal particles thus aiming to interpret the coking behavior of coals (Hardgrove 1932). HGI correlates to VM content, and the relationship is empirically specified for most of the hard coals and given with VM from 10% to 38% (db) by Eqs. (11.6) and (11.7):

For the execution of each test, further coal property parameters, particle size distribution and moisture content, as well as the height of fall of the stamp and the number of stamping steps were kept constant, so that the only parameter varied was the coal rank characterized by HGI.

The obtained data of each test was analyzed as described above to calculate the stampability factor K. A higher value for the HGI is equivalent to a lower resistance to stamping, i.e., a better stampability. The determined values of the stampability factor K are plotted against HGI in Fig.11.12.

## technical information - smc testing

The SMC Test was developed to provide a range of useful comminution parameters through highly controlled breakage of rock samples. Drill core, even quartered small diameter core is suitable. Only relatively small quantities of sample are required and can be re-used to conduct Bond ball work index tests.

The results from conducting the SMC Test are used to determine the so-called drop-weight index (DWi) which is a measure of the strength of the rock as well as the comminution indices Mia, Mih and Mic. In conjunction with the Bond ball mill work index they can be used to accurately predict the overall specific energy requirements of circuits containing:

For detailed information about how this can be done click here for the document Using the SMC Test to Predict Comminution Circuit Performance (PDF, 99KB) which describes the equations and/or click here to use our on-line specific energy calculator. This easy-to-use calculator is based on equations which have been benchmarked against large and varied data bases of operating circuits and equipment, ensuring that the resultant predictions should match operating experience as shown in Figures 1-3 .

The SMC Test also generates the JK rock breakage parameters A, b and ta as well as the JK crusher models t10-Ecs matrix, all of which are generated as part of the standard report output from the test. These values can be used to simulate crushing and grinding circuits using JKTechs simulator JKSimMet.

The SMC Test is a precision test that uses either crushed rock pieces that are very closely sized (so-called crush and select method) (Figure 4) or particles that are cut to similar size from drill core using a diamond saw (Figure 5). The latter approach (so-called cut-core method) is used when limited drill core sample is available. Almost any drill core size is suitable, even core that has been quartered (slivered). The chosen particles are broken using a closely controlled range of impact energies. The high degree of control imposed on both the size of particles and the energies used to break them means that the test is largely free of the repeatability problems which plague tumbling mill rock characterisation tests. Such tests usually suffer from variations in feed size, which are not closely controlled, as well as energy input, which although is often assumed to be constant is often highly variable.

The question is sometimes asked as to whether there is any difference in SMC Test results between using cut core or crushed core samples. This was considered during the early development of the SMC Test and testing protocols were tailored to ensure that no differences resulted. To confirm this experiments were carried out in which large lumps of ore were cored and the remains of the lumps then crushed. Test pieces were then prepared from the drill core using the cut core method whilst the crush and select method was used on the crushed lump material. Both types of sample were then broken independently in the drop-weight tester to determine if there were differences in the results. The raw data from these tests are shown in Figure 6. No significant difference between the two data sets was found.

The amount of sample that is required depends on what sources are available to provide the rock samples eg, is it drill core, what size of core is it, is it whole, halved or quartered as well as the size fraction chosen to do the SMC Test and whether the sample is going to be prepared by crushing or cutting. These factors are best discussed with your local metallurgical laboratory when you are at the planning stage. However, in the majority of cases 15-20 kilograms is more than enough to conduct a single test if feed preparation is being carried out by crushing the drill core first of all (crush-and-select method). However, if the sample is to be prepared by cutting with a diamond saw (cut-core method) as little as 5kg is normally sufficient. Also remember that the products from the SMC Test can be re-used for Bond ball work index testing, the SMC Test being effectively used as a feed preparation step for the Bond ball work index test. As a general rule Bond ball work index tests are recommended when SMC Tests are carried out as this provides valuable additional information on the way that much finer particles break.

Contrary to what some published literature claims, the number of samples required is related to the variability of the deposit and not the type of comminution test used. Also the end use will influence the number of samples required. Hence if samples are required for a pre-feasibility study the number will be relatively low, whilst if samples are required for the development of a Geometallurgical model that has the ability to accurately forecast daily grinding circuit throughput, the number required will be at least an order of magnitude higher. In all cases a staged approach to sample selection and laboratory testwork is recommended to ensure that costs are kept to a minimum. Each stage should be designed to build on the knowledge gained from preceding ones, particularly concerning variability, both spatially within the pit as well as in terms of absolute hardness values. For information on how to determine what the appropriate number of samples should be, click here to download the paper How to Formulate an Effective Ore Characterization Program

Formerly called Calibration, a more appropriate term is particle Size Scaling. NO SIZE SCALING is required to generate Mia, Mih and Mic parameters from a SMC Test as they are fixed functions of the DWi, which in turn is produced as a standard output from the SMC Test.

Where the SMC Test is used to estimate values of A and b (used in the JK AG/SAG mill model) a so-called Size Scaling factor may be required. The relationship between the DWi and the JK rock breakage parameters A and b makes use of the size-by-size nature of rock strength that is often apparent from the results of well controlled tests on different size fractions. This is illustrated for a range of different rock types in Figure 8, which shows how the gradients of the plots of normalised values of A*b vary with particle size. In the case of a conventional drop-weight test the A*b values for each particle size are effectively averaged and a mean value of A and b is reported. The SMC Test uses a single size and makes use of relationships such as that shown in Figure 8 to predict the A and b of the particle size that has the same value as the mean for a full drop-weight test through the use of a size scaling factor. The average particle size of a full drop-weight test is approximately 30mm and where the SMC Test is carried out on particles similar to this size the size scaling factor approaches unity. In cases where the particle size used in the SMC Test is not 30mm, eg where only relatively small diameter drill core is available or the drill core has already been slivered into half or quarter core, then a size scaling factor will need to be applied.

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