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By E. M. Wise

THOUGH known as the platinum-group metal- the sextuplet, platinum, palladium, iridium. rhodium, osmium, ruthenium, might well be called the American metals or perhaps Pan-American metals, as the ore containing them eras first discovered in South America, and in recent years the largest production has come from North America. The South American alluvial deposits, which also, contain gold, are in active production, as are the alluvial platinum deposits of Alaska, but the main production is from the copper-nickel ores of Sudbury, Ontario. The Russian deposits in the Urals were the source of most of the worlds platinum for about 100 years, but since 1934 have assumed second place. For a time considerable amounts of platinum came from several rather rich deposits in South Africa but these were exhausted quickly. However, a tremendous reserve of lower-grade ore exists there, which could be worked profitably if the price of platinum were considerably higher.

Jan 1, 1942

By Gilbert L. Eggert

The effects of reactor irradiation on age hardening in a Cu-2 wt pct Be alloy have been investigated in two neutron-energy ranges. During some of the irradiations, specimens were exposed to high-energy fission neutrons which passed through a boron-10 shield; during other bombardments, the boron-10 was omitted and specimens were irradiated in a reactor thermal column, or the fuel configuration was modified to increase the thermal-neutron flux at a location in the core. Differences in length, X-ray diffraction angle, and hardness indicated that low-energy neutron irradiation made a small contribution to changes in the measured properties. The influence of defects generated by knock-om with less than the Brinkman critical energy or by prompt (n,y) recoil is suggested as the cause of this effect. EXPERIMENTS designed to test the effects of reactor irradiation on the properties of age-hardening alloys have to the present time employed fluxes un-differentiated according to neutron energy. Billing-ton and Siegell in 1950 were the first to perform such an experiment and to compare the results with those obtained by thermal aging. These authors concluded that precipitation in a Cu-2 wt pct Be alloy was not enhanced by the radiation. However, 4 years later, Murray and Taylor,2 working with specimens of the same Cu-2 wt pct Be alloy, reported that an increase in density, hardness, and resistivity which followed reactor irradiation resembled changes associated with thermal aging at about 100°C. (It should be noted that these authors reported that higher aging temperatures caused a decrease in resistivity.) cupp3 reported hardness results similar to those of Murray and Taylor. Koppenaal4 showed that irradiation-enhanced precipitation and irradiation defect hardening were not simply additive but that they acted separately in an irradiated alloy. Kernohan et a1.,5 working with Ni-Be, offered evidence that beryllium was removed from solid solution by reactor irradiation and suggested that the mechanism was one of increased diffusion through the use of vacancies in the structure. piercy6 also found that more precipitate particles were formed by fission-neutron irradiation of a Cu-Co alloy which contained a small amount of particles before irradiation. He proposed that the additional particles were caused by an appreciable amount of diffusion during the final cooling of displacement spikes. The retrogression of extant precipitate particles found by Murray and Taylor2 and Boltax7 as well as piercy6 indicated that nuclei were dissolved within a high-energy spike region and, upon resolidification of the perturbed volume, new nuclei precipitate and grow by enhanced diffusion. Bleiberg et al.' and Konobeevski et al.' explained phase changes obtained by irradiating fissionable alloys in terms of retrogression and re-precipitation in regions of high-energy spikes. Parsons and Balluffi10 used electron microscopy to verify that crystallization did indeed occur in displacement spikes which were formed in amorphous germanium by fast-neutron irradiation. The concept that low-energy neutron irradiation might cause structure-sensitive property changes in semiconductors by atomic recoil associated with the (n, y) reaction was first proposed by Schweinler" and walker." This thesis was tested by Cleland and crawford" and found to be valid. Coltman et al.14 exposed a variety of pure metals to thermal-neutron irradiation at cryogenic temperatures and attributed the resulting resistivity changes to the action of defects created by (n, y) recoil. The present work was undertaken to extend the concepts of Schweinler," walker," and Coltman et al.I4 to the case of an age-hardening alloy and to determine whether or not the action of defects produced by thermal-neutron irradiation would en-hance age hardening. EXPERIMENTAL PROCEDURE Omeea West Reactor Irradiations.* Thermal-

Jan 1, 1965

By P. P. Gregory

ESTIMATION of re- serves in prorated sand fields has been discussed by S. A. Judson, H. D. Easton, Jr., and W. A. Schaeffer, Jr., in a paper that appears in Vol. 114 (1935), of the A.I.M.E. TRANSACTIONS. The purpose of this present paper is to extend the investigation to prorated flush limestone fields and to outline a method of procedure in determining the amount of oil originally in place and the amount of oil that can be recovered from such fields. Although it is possible to use the decline curves to estimate the amount of recoverable oil in certain limestone fields that are allowed to produce the oil at or near capacity production, such a method cannot be used in fields under severely restricted production rates. The problem is exceedingly complex due to fluctuation in the allowable rates of oil production over a wide range and at frequent intervals, and especially because of the heterogeneous nature of the reservoir itself. A well completed in a large solution cavity, capable of producing great quantities of oil per day, may be offset by a well of very limited capacity. Also, it is di5cult to arrive at an average figure for porosity in a given limestone reservoir due to abrupt variations in the character of limestone, in many instances even between two ad- joining wells. It is a safe statement that uniformity of porosity and permeability in limestone reservoirs .is nonexistent. It is evident, therefore, that the method of estimation of re- serves on the basis of per-acre content can give only approximate results.

Jan 1, 1935

By R. W. Gurry, L. S. Darken

The solubility of cementite in austenite is computed by thermodynamic methods from the observed solubility of graphite. It is found that the solubility of cementite is greater than that of graphite in the entire austenite temperature range. Thus the prior discrepancy between this portion of the phase diagram and the observed metastability of cementite is partially resolved. FOR several years it has been apparent that a serious discrepancy exists in certain observations pertaining to the Fe-C system. Direct experimental data indicate: 1-That the curve representing the solubility of graphite in austenite'" crosses that for cementite4 at about 945°C as shown in Fig. 1. 2— That graphite and austenite saturated therewith are stable1 relative to cementite at all temperatures between the eutectoid, about 723°C, and the eutectic, about 1153°C. These observations are mutually incompatible in view of a well-known consequence of the second law of thermodynamics, namely, that the stable phase (presumably graphite at all temperatures) has a lower solubility than the metastable phase (presumably cementite). Hence either 1 or 2 above must be in error; either the measured solubility of graphite or of cementite, or the observation as to the stability of graphite, is in error. Our present purposes are: 1—To collate the available data on cementite and prepare a table of HO — HO° and of (Fa— HO°)/T. This tabular method seems the best and most convenient way to represent the heat and free energy of formation. 2—To determine whether the measured solubilities of cementite and graphite in austenite are in accord with the above table and the measured thermodynamic properties of austenite, in order to resolve, if possible, the existing paradox of these two solubilities. Enthalpy of Cementite The enthalpy of cementite at 273 °K, H279 — HOcem, was computed by integration of the low temperature heat capacity as measured by Seltz, McDonald, and Wells" 68" to 298 °K) and as given by the Debye functions recommended by them (0" to 68°K); the result is included in Table I. At higher temperature HCem — H,,"" is given by Kelley who used the data of Naeser7 (298" to 1037°K) and of Umino8,9 (273° to 1923°K). We have recalculated Hcem — H273 cem from the data of UminoO in the following way. In the equilibrated alloy at temperature the weight fraction of cementite Pct c -Pct cr is PctC -PctCr designated f(C), and the

Jan 1, 1952

By M. M. Rao, P. G. Winchell, R. J. Russell

A detailed analysis has been used to correlate the avalable thermodynami c da In on iron-ric11 Fe-Ni (41loys in the body-centered a, and the face-centered y phases. The inp ut injovrt~ation required by the a am lysis consists of tile heat of the a, to y transformation at 1123°K, the heat of mixing of y at 1123°K, tile specific heats of both a, and y (decomposed into Debye, electronic, and magnelic contributions), and the positional and magnetic entropies of both a, and y. PYLrrzary oirt.bul of the analysis is the free energy of each phase as a Junction of temperature and composition. FOr Fe-Ni alloys the adjustment of twenty-one parunletevs allows the accurate reproduction of enthalpy data for both phases, the heat of mixing- for y, the a/y equilibrinm compositions in the Fe-Ni phase diagram, and the usually accepted enthalpy and entropy of a and y iron. Also obtaitzed are fuirly accurate reproductions of electronic specific heat coefficients predicted from electron- to-atom ratio interpolations and To temperatuves eslimanted from transformation studies. Also considered in the calculation are the elastic constants, Curie temperatures, saturation magnlelizalion values, and indications of short- range order. The measnred activity coefficient of iron in anstetnite could not be accurately matched. The major source of uncertainty in the model is thought to be the magnetic contribution to the entropy of austenite. Calcrclations were extended to Fe-ni-C alloys using available carbon-acliuity data. The free energy of transformation from martensile to r at fixed composition was obtained as a function of nickel and carbon contetzt, and binary thetal-carbon phase diagrams Were estimaled as a function of nickel contenl. We needed thermodynamic information for Fe-Ni and Fe-Ni-C alloys at temperatures between 400" and 800°K in order to interpret the results of several phase-transformation studies. The most recent theoretical analysis of the thermodynamics of Fe-Nialloysl was based on a lumped-parameter, regular-solution model and used the Fe-Ni phase diagram available at that time.' Because significant difficulties were encountered in attempting to extend this analysis to Fe-Ni-C alloys, and because more experimental data has become available recently, we decided to attempt a new correlation. The analysis used here allows the extrapolation of measured and calculated thermodynamic values to lower temperatures and the extension of the thermodynamic analysis to Fe-Ni-C alloys. The structure of this correlation consists of expressions containing a total of twenty-one adjustable parameters. The forms of these expressions are discussed in the next section; however, a less formal summary is presented now to provide a physical overview of the model and to indicate its connections to prior work. Primary input information for this correlation is described below. The measured values of the specific heat of iron3 and its analysis475 and theoretical estimates of the specific heat of y* iron at low tempera- loys at 1123°K has been measured by solution calori-metry.7 The enthalpies of both a, and y have been measured most recently by Scheil and Saftig,B who used a dry-ice calorimeter. These data are used to formulate expressions for the heat of mixing of y and the heat of transformation of a, to y both at 1123°K. The a and y solubility limits in the Fe-Ni phase diagram have been redetermined recently9 by methods which assure a close approach to equilibrium. The a to y transformation can occur without change in nickel content or free energy at a temperature called TO; TO temperatures for various nickel contents have been estimated from transformation studies. The computed free-energy functions are adjusted by means of the parameters to fit both the phase diagram and the TO temperatures. Primary emphasis is placed on the phase diagram fit. Iron activities have been measured in Fe-Ni austenite10 between 973" and 1173°K. These activity data, which cannot be fitted accurately with our model, indicate that iron is substantially ideal in iron-rich austenite. Secondary, but helpful, information for the present correlation is provided by neutron diffraction and saturation magnetization measurements on both body-centered and face-centered Fe-Ni alloys."-'3 These results were interpreted11 in terms of composition-insensitive, localized magnetic moments on both iron

Jan 1, 1968

By David R. Maneval, H. B. Charmbury

At the turn of the century, iron and coal were the keys to industrial prosperity. At that time, Pennsylvania was the leading mineral producer in the Country, producing 200,000,000 tons of coal in a typical post-World War I year. Close to 90% of the Nation's energy was generated by coal. Today, coal accounts for but 22% of the energy generated, leading to a drastic reduction in Pennsylvania coal output. In contrast with the post-World War I days, 1964 figures showed a drop in production to 93,846,480 net tons. This huge decrease has created a serious economic problem in a state so much dependent on the coal industry. Along with the reduction in coal production, employment has dropped considerably through these years (the number of miners alone has gone from 280,011 to 43,149), thus causing a heavy burden of supporting the unemployed on the taxpayers. As far back as 1939, coal operators and economists saw the problems on the horizon and the need for improvement. Realizing that research was a way to improve and stabilize the industry they demanded legislative action. The Legislature reciprocated with a small grant to The Pennsylvania State University with the provision that each State research dollar be backed with one from the industry. This initial grant laid the foundation for the further growth in coal research.

Jan 3, 1966

By B. W. Frazier

Jan 1, 1882

By Bunting S. Crocker

PEBBLE grinding at Lake Shore is not a temporary wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct - 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 1 1/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to, $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change, back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1 1/4-in. steel ball charge. To change pebble size it is necessary only to change two sets-of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean. out the customary supply of grinding balls kept on hand. Also it has not always been ' possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant should be designed to use as many units-or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, .2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant. and an extremely efficient-one. In this plant the ratio of ball mills to tube mills was 1: 2. When the much cheaper pebble mills were substituted for the tube mills, this. ratio -was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was. proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on -.8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.1 This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without .grate discharges-both with 1 1/4-in. and 3/4-in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report' was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3 1/2-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are, lengthy and may be covered in a separate 'report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The -8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands .from the bowl sent to the primary pebble mills, see Fig: 1. In the pebble mill circuit the ore is ground to 90 pct -325 mesh (24 pct + 28 microns) In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO2, 80 pct insoluble. Its grindability at different. meshes has been shown near the top of 'the list in F. C. Bond's grindability tests' Lake Shore is not shown, but an adjacent-mine with identical ore, Wright-Hargreaves, is. Reasons for the-Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1952

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By Elmars Ence, Harold Margolin

The titanium-aluminum system has been investigated in the composition region 0 to 34 pct Al in the temperature range 800" to 1450°C. The phases encountered in this region were: a,ß, TiAl3, The reactions observed are as follows: 13 ; L - 6T. A1, 0 (15 pct Al) + 6 (18 pct Al) - yTi Al (16.5 pct Al) at -1170°C andP (7 pctpct Al) + y (9.2 pet Al) - ah(9 pct Al) at - 1060°C. A eutectoid reaction: y — a +d belm 700°C is postulated. THE Ti-A1 system has been the sub.ect of a number of investigations. Several of these1-3 have indicated extensive solubility of A1 in both a and ß titanium. Ence and Margolin, as a result of work on the Ti-A1-0 system:'5 have shown that the a solubility ofalumi-num is considerably restricted, and found that a compound, Ti2A1 existed in the region formerly designated as Q a This structure was identified as hexagonal with c/a = 0.803. Pietrokowsky confirmed the existence of this phase7 and pointed out that the structure was isomorphous with Ti3Sn.8 Although the existence of Ti,A1 has been confirmed by others,9-12 there has been considerable disagreement regarding the composition and mode of formation. A detailed reinvestigation of the Ti-A1 system has been published by Sagel et al.13 The tentative diagram indicated two phases, a, and e, occurring in the former all a region. The a, phase has the same structure as Ti2A1 and was indicated as having a broad region of solubility. Epsilon, a tetragonal phase, was shown to form within the a, solubility region. AS a result of unpublished work,14 Ence and Mugy3 olin disagreed with the interpretation of Sage1 et al., although some common features existed. The present study was conducted in an attempt to clarify prevailing uncertainties. EXPERIMENTAL PROCEDURE Alloy Preparation—Alloys were prepared from iodide titanium (99.9 pct Ti with less than 0.01 pct Zr) or Bureau of Mines electrolytic titanium (BHN-73) and high purity aluminum (99.99 pct). Alloys were arc-melted in argon or helium atmosphere as 15-g buttons in the composition range up to 34 pct Al. Depending on need, alloys were made in increments varying from 0.25 to 2 pct Al. Up to five 15-g buttons of some compositions were made. Weight losses during melting did not exceed about 1 pct of the charge and chemical analysis revealed that both titanium and aluminum were evaporated. In general, nominal and analyzed compositions agreed within 0.5 pct and therefore nominal compositions were used in plotting the data. Unavoidably, some overlap of composition occurred, particularly where small increments of A1 were used. The small increments were primarily used to obtain alloys which would fall into the narrow a -, r region and this purpose was achieved. In general, the data obtained from Bureau of Mines base titanium and iodide titanium alloys were in agreement. Forging—Both as-cast and forged samples were employed. Samples were heated for hand forging by an oxygen-hydrogen torch to the region 1200" to 1300°C. A titanium anvil and cover plate were used to prevent any iron pick-up and consequent formation of a low melting point compound. Frequent reheating kept the temperature in the desired range. Under these conditions, alloys containing up to 15.5 pct could be successfully forged. To check on the depth of contamination, forged alloys were heated slightlybelow the ß transformation temperature. Metal was removed to a depth below the contaminated surface so revealed, and the subsequent results obtained from these samples agreed closely with those secured from as-cast material. Heat Treatment—Prior to heat treatment, all alloys were vacuum annealed to remove hydrogen. Samples

Jan 1, 1962

By J. L. Schroeder, W. G. Talman

EXPERIENCE has shown that certain known natural conditions and other indefinite characteristics combine to make a mining area vulnerable to mountain bumps. Some of the known conditions are heavy overburden, an overlying stratum of strong non-elastic rock, a structurally strong coal seam that does not crush easily and yet is the weakest stratum in the series, and a floor stratum of more than ordinary firmness. The indefinite characteristics could include resistance of top or bottom of a coal seam to lateral flow, effect of immediate top and bottom strata, nature of the coal seam, and other factors about which very little is known. It is in this area that exhaustive study is needed to achieve positive control of the dreaded mountain bump hazard. In the Gary district, West Virginia, two of the five operating mines have areas known to be vulnerable to mountain bumps.These are the No. 2 and 6 mines, which are working the Pocahontas No. 4 coal seam. In this area the seam is overlain at a zero to 10-ft interval by the massive Eckman sandstone, which is about 100 ft thick in the vulnerable areas. Total overburden, consisting of shale, coal seams, and sandstones, ranges from 700 to 1400 ft where bumps have occurred (Fig. 1). The coal seam is 4 to 8 ft thick and has a weak columnar structure of very soft and spongy or woody coal. The immediate top is either the Eckman sandstone or hard shale, both of which provide generally good roof conditions for mining operations. The bottom ranges from medium to hard gray shale. The incidence of bumps in the Gary district dates from about 1930 and can be correlated to some degree with the increased rate of extraction attained with mechanical mining. During early years of operation at the No. 2 and 6 mines, most of the work was development and under comparatively light cover. As overburden became heavier, together with an increasing percentage of coal coming from pillar extraction, mountain bumps began to occur. Before 1945 they were much less severe than in later years, and because there were no injuries or fatalities, no records were kept. But by 1951 the incidence of bumps had increased to such extent that it became evident a mining system would have to be designed to eliminate or at least minimize these occurrences. Drawing on experience with modified versions of old systems and certain characteristics peculiar to the Pocahontas No. 4 coal seam, the district mining engineer, Martin Hayduk, outlined the essential factors for bump control in the Gary district. These control factors indicated: 1) The highly stressed area, which lies within the limits of convergence effects, moves at approximately the same rate as the retreating pillar line. 2) There is a decided softening of the periphery of blocks which is greatest and deepest on the gob side. 3) Blocks less than 45 ft wide never have been known to bump in the Pocahontas seams. Fig. 2 shows the mining plan based on these observations—a thin-pillar system with a minimum of secondary development. It was recognized from the start that the Hayduk system had certain limitations. When pillar work backed up to an inevitable surrounded block it was no longer applicable; also, burnt bottom and/or unknown factors reduced the system's effectiveness in certain headings. Since there was no way to eliminate the loaded block, it was decided to experiment with auger drilling to attain controlled release of this stored-up energy. The decision to auger was made after consideration of three separate plants, which were: 1) To cut the solid block into pillars of maximum size by driving away from the fracture line and at such a distance outby that development

Jan 1, 1959

By Jerry Tien

Mine regulators are normally used for proper air distribution in underground mines. They are deliberately introduced resistance in the regulated airway, and by altering sizes, they can distribute specific air quantities to regulated splits and free splits (4, 5). The enactment of the Federal Coal Mine Health and Safety Act of 1969, the Federal Mine Safety and Health Amendments Act of 1977, and the regulations adopted by the Mining Enforcement and Safety Administration (now called the Mining Safety and Health ministration) prompted greater discussions of optimal mine regulator locations because of the Act's stringent ventilation requirements (2, 3). For example, by requiring the separation of the air used in the belt haulage entries from the intake and return airways, the regulations required the creation of a third type of airway---the neutral airway, as opposed to the intake and return airways, adding to the complexity of a mine's ventilation network. It is the intent of this report to show that panel regulators should be located in intake airways at the neck of a panel for a blowing fan system and in return airways at the neck of a panel for an exhausting fan system in order to achieve optimal air control. A computer simulation model is used to determine optimal mine regulator location.

Jan 1, 1983

By A. Rosen, P. Wynblatt, J. E. Dorn

The temperature, strain rate, and structure dependence of the flow stress in a poly crystalline Fe-2 pct Mn alloy was investigated between 77" and 370°K. It was possible to identify the low-temperature data from 77° to 160°K with Dorn and Rajnak's theory of the Peierls mechanism of plastic deformation. Above 160°K one or more thermally activated mechanisms must he operative; however, these were not identified in this investigation. The mechanistic basis for the rapid decrease in the flow stress of bcc metals with an increase in temperature over the low-temperature range is yet being debated. Various dislocation mechanisms have been suggested in the past but the agreement between the various proposed theories and the experimental results have not been definitive. In the present paper, the authors will demonstrate that the plastic behavior of polycrystalline iron containing 2 wt pct Mn plus interstitials agrees excellently with predictions based on the Peierls mechanism from 77" to 160°K whereas from 160" to about 370°K other as yet unidentified thermally activated mechanisms take over. Undoubtedly, the overlap of these mechanisms with the Peierls mechanism contributed to the difficulties of rationalizing the low-temperature behavior of iron and its alloys. I) EXPERIMENTAL TECHNIQUES AND RESULTS Although extensive data are now available on the effect of temperature and strain rate on the flow stress of various bcc metals and alloys, it was, nevertheless, deemed necessary for the objectives of the present study to obtain highly accurate data of the type most pertinent for checking the theories on a series of well-documented work-hardened states. Iron containing 2 wt pct Mn was used in this investigation in order to obtain a lower ductile to brittle transition temperature than that for iron.' The additional elements present were 0.004 pct C, 0.05 pct 0, 0.004 pct S, 0.003 pct P, 0.006 pct N, and 0.001 pct Si. The as-received 3/8 by 3/4 in. bars were cold-rolled to 0.100-in. thickness, recrystallized under argon for 30 min at 1073"~, further cold-rolled to 0.063-in. thickness, and finally recrystallized under argon for 30 min at 1000°K. This treatment gave a stable reproducible equiaxed grain structure having an average grain diameter of about 90 p. Prior to the final recrystallization treatment, the sheet was machined into tensile specimens 1/4 in. wide having a 1.625-in.-long gage section. All specimens were tested in controlled-temperature baths on an Instron Testing Machine. For convenience in the theoretical analyses the data will be given in terms of the shear stress for flow, 7, the shear strain rate, f, and the shear strain, y. A series of three standard strain-hardened states were obtained by prestraining specimens at 300°K and at a shear strain rate of 7.68 x lo-' per sec to three stress levels of 6.89 X l08, 10.3 X lo8, and 14.5 x 10' dynes per sq cm, respectively. Immediately following this prestrain the temperature bath was changed and each specimen was pulled in tension at either f = 7.68 X lo-' or f = 3.06 X 10"3 per sec. In terms of the accuracy that was achieved, the values obtained did not deviate more than *0.1 X 10' dynes per sq cm and * 1°K from the reported values of stress and temperature and not more than *5 pct in the reported values of the strain rate. The tensile stresses were converted to shear stresses using a factor of 1/2. The shear stress for the initiation of flow at the test temperature was determined by taking a y = 6.16 x 10'3 offset. from the modulus line at which point good accuracy could be achieved in determining y and the associated flow stress 7, with only a negligible increase in the pre-

Jan 1, 1965

By Owen Thornton

Recently equations have been presented by Rose and Bruce' and by Rose², showing how the relative permeability of a reservoir rock may be determined from the capillary character of the rock. In particular, equations were developed to show the relationship between capillary pressure and the effective permeability to the wetting phase in a poly-phase flow system. The equations are as follows (nomenclature the same as in the Rose and Bruce paper) Rose and Bruce assume that the average length of path (L,), which the wetting fluid follows in flowing through a porous media is independent of the saturation to the wetting phase, so that 1 is a constant dependent only on the k rock, and the permeability to the wetting phase is a function of the saturation and of the capillnry pressure: In measuring the flow of ekctric current through partially saturated core samples, it has been observed that the electrical resistance usually is not a simple linear function of saturation. This fact indicates that the average length of path in the flow of electric current is not independent of saturation, but rather that the tortuosity of the path depends upon the average saturation to the conducting fluid as well as upon the characteristics of the rock. itself. Inasmuch as the flow of fluid and the flow of electric current in many respects are analogous, it may be indicated further that the length of path for fluid also may not be independent of saturation. The following relationship can be derived to express the resistance to flow of electric current through a core sample partially saturated with conductive wetting phase, the non-wetting phase being non-conductive: where L is the average length of path followed by the current in flowing through a length L of partially saturated core, where L is the length of current path when the core is fully saturated, and where Rs and R, are the specific resistivities of the partially saturated and saturated core sample, respectively. It will be noted that R., the specific resistivity of the core sample when S, = 1.0, is equal to the "formation resistivity factor"' multiplied by the specific resistivity of the saturating fluid. If it be assumed that the average lengths of path for fluid flow and for the flow of electric current are the same, then equation (5) may be combined with equations (l), (2) and (3) to give the following relationship: The above equation suggests including a correction for tortuosity when calculating the relative permeability of a reservoir rock to a wetting phase. The correction factor may be obtained from the resistivity-saturation relationship. For instance, the following calcu- lations are obtained for the unconsoli-dated sands described by Leverett Kw Kw Sw R°/Rs Pt/Pc¹ (Calc.) (Obs.) 0.30 0.08 0.88 0.02 0.0 0.40 0.16 0.94 0.06 0.04 0.50 0.27 0.97 0.14 0.11 0.60 0.40 0.98 0.26 0.21 0.70 0.53 0.98 0.39 0.35 0.80 0.68 0.99 0.57 0.54 0.90 0.84 0.99 0.77 0.76 1.00 1.00 1.00 1.00 1.00 It will be noted that the agreement between the calculated points and the observed data is rather good. Further investigation of the above method for obtaining relative permeabilities may be merited. For many sands within specified ranges of saturations the following relationship has been found to hold approximately': — =SW"......(7) Rs The exponent n in the above equation has been found to equal two for many sands so that equation (6) reduces to the following: The writer acknowledges the permission of The Texas Co. to submit this note for publication. 1. Walter Rose and W. A. Bruce—Jnl. of Pet. Tech., Vol. 1, No. 5, p. 127 (1949). 2. Walter Rose—Jnl. of Petr. Tech., Vol. 1, No. 5, p. 111 (1949). 3. G. E. Archie— -Trans. AIME, 146, 54 (1942). 4. M. C. Leverett— Trans. AIME, 142, 152 (1941). 5. M. C. Leverett—Trans. AIME, 132, 149 (1939).

Jan 1, 1949

By Owen Thornton

Recently equations have been presented by Rose and Bruce' and by Rose², showing how the relative permeability of a reservoir rock may be determined from the capillary character of the rock. In particular, equations were developed to show the relationship between capillary pressure and the effective permeability to the wetting phase in a poly-phase flow system. The equations are as follows (nomenclature the same as in the Rose and Bruce paper) Rose and Bruce assume that the average length of path (L,), which the wetting fluid follows in flowing through a porous media is independent of the saturation to the wetting phase, so that 1 is a constant dependent only on the k rock, and the permeability to the wetting phase is a function of the saturation and of the capillnry pressure: In measuring the flow of ekctric current through partially saturated core samples, it has been observed that the electrical resistance usually is not a simple linear function of saturation. This fact indicates that the average length of path in the flow of electric current is not independent of saturation, but rather that the tortuosity of the path depends upon the average saturation to the conducting fluid as well as upon the characteristics of the rock. itself. Inasmuch as the flow of fluid and the flow of electric current in many respects are analogous, it may be indicated further that the length of path for fluid also may not be independent of saturation. The following relationship can be derived to express the resistance to flow of electric current through a core sample partially saturated with conductive wetting phase, the non-wetting phase being non-conductive: where L is the average length of path followed by the current in flowing through a length L of partially saturated core, where L is the length of current path when the core is fully saturated, and where Rs and R, are the specific resistivities of the partially saturated and saturated core sample, respectively. It will be noted that R., the specific resistivity of the core sample when S, = 1.0, is equal to the "formation resistivity factor"' multiplied by the specific resistivity of the saturating fluid. If it be assumed that the average lengths of path for fluid flow and for the flow of electric current are the same, then equation (5) may be combined with equations (l), (2) and (3) to give the following relationship: The above equation suggests including a correction for tortuosity when calculating the relative permeability of a reservoir rock to a wetting phase. The correction factor may be obtained from the resistivity-saturation relationship. For instance, the following calcu- lations are obtained for the unconsoli-dated sands described by Leverett Kw Kw Sw R°/Rs Pt/Pc¹ (Calc.) (Obs.) 0.30 0.08 0.88 0.02 0.0 0.40 0.16 0.94 0.06 0.04 0.50 0.27 0.97 0.14 0.11 0.60 0.40 0.98 0.26 0.21 0.70 0.53 0.98 0.39 0.35 0.80 0.68 0.99 0.57 0.54 0.90 0.84 0.99 0.77 0.76 1.00 1.00 1.00 1.00 1.00 It will be noted that the agreement between the calculated points and the observed data is rather good. Further investigation of the above method for obtaining relative permeabilities may be merited. For many sands within specified ranges of saturations the following relationship has been found to hold approximately': — =SW"......(7) Rs The exponent n in the above equation has been found to equal two for many sands so that equation (6) reduces to the following: The writer acknowledges the permission of The Texas Co. to submit this note for publication. 1. Walter Rose and W. A. Bruce—Jnl. of Pet. Tech., Vol. 1, No. 5, p. 127 (1949). 2. Walter Rose—Jnl. of Petr. Tech., Vol. 1, No. 5, p. 111 (1949). 3. G. E. Archie— -Trans. AIME, 146, 54 (1942). 4. M. C. Leverett— Trans. AIME, 142, 152 (1941). 5. M. C. Leverett—Trans. AIME, 132, 149 (1939).

Jan 1, 1949

The Library of the above-named Societies is open from 9 A.M. to 10 P. M. except on holidays. It contains about 70,000 volumes and 90,000 pamphlets, including sets of technical periodicals and publications of scientific and technical societies. Members of the Institute, with few exceptions, are forced to spend a portion of their time in localities isolated from sources of information. To these the Library, through its Library Service Bureau, can render valuable service through correspondence; letters requesting information will receive especial attention. The Library is prepared to furnish references and photographic copies of articles on mining and metallurgical subjects; to determine the existence of mining maps, and to furnish general information on the geology and mineral resources of all countries. All communications should be made as definite as possible so that the information received may be what is desired and not include collateral matter which may not be of interest. The tine spent in searching for such collateral matter will be saved, and the information will be sent more promptly and in more usable shape.

Jan 10, 1917

By AIME AIME

A DOCUMENT of progress and of great interest to engineers is the report of the Military Affairs Committee of 'the Engineering Council, which has just been accepted and sent to the secretary of War. It provides .a plan whereby civilian engineers holding commissions in the Reserve Corps may; with their con- sent, be called into service of the army during peace at any time, either with troops in the field or in a consulting capacity on any government engineering work. The Military Affairs Committee finds ample authority in the military bill passed by Congress on June 4, 1920, for this arrangement and also for extending the privilege to military engineer, officer-s of the regular army to engage in civilian work during times of peace. It is hoped that the Secretary of War will adopt the suggestions and put the program outlined into effect. This is a constructive piece of work comparable to the efforts of the military Committee before the war which developed so much in the way of preparedness, particularly in the program of military lectures which, held in the auditorium in the Engineering Societies Building, became so popular that each lecture had

Jan 1, 1920

By K. P. Gupta, A. K. Pal, L. Chandrasekaran, U. P. Singh

The Mn-Sn binary system, investigated at the high-manganese end and between 500° and 1000° C, shows four phases at temperatures below 727"C, namely the u Mn, the p Mn, the Mn3 Sn, and the Mn, Sn phases, while at higher temperatures only the last three phases remain stable. The solubility of tin in a Mn is very small and the maximum solubility of tin in P Mn phase appears to be abollt 10 at. pct Sn. The solubility range of the Mn,Sn and the Mn, Sn phases is 23.0 to 26.0 and 37.0 to 40.0 at. pct Sn, respectively. The lattice parameter of the 13 Mn phase increases with increasing tin content. The Mn, Sn thase is hexagonal with and appears to be basically of the Ni3 Sn type structure except that for quite a few X-ray diffraction lines the calculated and observed relative intensities do not agree well. The Mn-Sn binary system has been studied by several investigators.''~ Their results indicate that three intermediate phases, namely, the Mn3Sn, Mn,Sn, and MnSn, phases, exist at high concentrations of tin. However, so far the proper phase equilibrium has not been established at the manganese-rich end and very little data is available for the composition range between pure manganese and Mn3Sn. Moreover, earlier investigators differ in their opinion about the exact composition range at which the Mn3Sn and Mn,Sn phases appear and some doubt has been cast by some investigators3 regarding the structure of Mn3Sn phase which has been reported to be isotypic with Ni3Sn (Mgscd type) structure. In this investigation an attempt has been made to establish the proper phase equilibrium between pure manganese and Mn + 50 at. pct Sn composition in the temperature range of 500" to 1000°C. PROCEDURE The raw materials used were from three different sources. For exploratory work five alloys containing 5, 10, 15, 20, and 25 at. pct Sn were prepared using 99.9 pct pure manganese and tin supplied by E. Merck & Co., Germany. The rest of the alloys and one more 25 at. pct Sn alloy were prepared using 99.9 pct Mn supplied by Gallard Schlesinger Chemical Mfg. Corp., U.S.A., and 99.999 pct Sn supplied by Semi Elements Inc., U.S.A. Weighed amounts of manganese and tin were melted in recrystallized alumina crucibles in an inert gas (argon) high-frequency induction melting furnace. By careful control of temperature and time of melting the losses were reduced to below 0.2 pct in all cases. Since the losses were very small, no attempt was made here to analyze the samples chemically. Alloys were wrapped in molybdenum foil and sealed in small evacuated fused silica capsules. The alloys were annealed at different temperatures, controlled within + l°C, for sufficiently long periods to attain proper phase equilibrium, and subsequently quenched in cold tap water. The annealing periods used at different temperatures were 15 days at 500°C, 7 days at 600°C, 5 days at 700" and 750°C, 3 days at 800°, 850°, and 900°C, 2 days at 968"C, and two alloys annealed at 968°C were reannealed for 10 hr at 1000° C. From each annealed specimen, a part was utilized for metallographic study while another piece was used for X-ray diffraction study. 1.0 pct HNO, solution and oxalic acid solutions of concentrations 0.05 to 1.0 pct were used for etching Mn-Sn alloys above and below the MnsSn composition, respectively. Since all alloys were brittle, X-ray specimens were prepared using the as-crushed -325 mesh alloy powders. Only one 25 at. pct Sn alloy powder was reannealed in an evacuated silica capsule at 800°C for 5 min and water-quenched. X-ray diffraction patterns . for the Mn3Sn phase with the as-crushed and the reannealed powders did not show appreciable change. Un-filtered iron radiation at 25 kv, 15 ma was used with either Norelco 114.6-mm-diam Debye Scherrer Camera (for phase identification) or Norelco 12-cm-diam symmetrical focusing camera (for lattice parameter determination of the 6 Mn phase). The estimated accuracy of lattice parameter determination for the focusing camera was * 0.001A. RESULTS AND DISCUSSIONS The results of metallographic and X-ray diffraction study made with different alloys are shown in Fig. 1. The variations in lattice parameter with composition for the p Mn and the Mn3Sn phases are given in Tables I and 11, respectively, and the lattice parameter as a function of composition for the p Mn phase is shown in Fig. 2. The lattice parameter of the p Mn phase increases with increasing tin content while for the Mn,Sn phase the data obtained from two two-phase alloys and one single-phase alloy indicate increase in a,, and decrease in c, parameters with increasing tin content. The results, Fig. 1, indicate that the solubility of

Jan 1, 1969

By V. E. Christensen

AT most smelting plants, forced draft, induced by high stacks or fans, is used to carry the gases away from the furnaces, roasters, or sintering plants. Gases moving under forced draft carry varying amounts of solids in the form of dust and fume. Where the weight of the solids is material, and the metal content high, it is usually profitable to treat the gases to recover these solids. Baghouses, Cottrell treaters, and settling chambers are the common forms of plants used to recover solids from gases. To check on the efficiency of these plants it is necessary to know the volume of gas leaving the plant and the amount of solids carried by the gas. As the volume of gas is equal to the product of the velocity of that gas and the cross-sectional area of the flue, the determination of the volume includes the measurement of the area of the cross section and the determination of the velocity. The velocity is generally measured by means of the Pitot tube and Ellison differential pressure gage, although low velocities are best determined by means of an anemometer. The dust concentration is determined by passing a measured volume of sample through a filter and weighing the dust caught. From this the dust concentration per unit of volume may be calculated. It is necessary to assume that the gas is homogeneous or that each cubic foot of the gas contains the same weight of dust as every other cubic foot. Having the volume of gas passing through a flue per unit of time, and the dust concentration of the gas per unit of volume, the amount of solids may easily be calculated.

Jan 1, 1935