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By S. A. Hoenig

In an earlier publication, the application of exoelectron technology to the monitoring of dry grinding in a ball mill was reported. Recently, the author has investigated the ball mill problem in more detail with the system shown in [Fig.1]. Here, some of the current generated in the ore-ball-water mixture was picked up by the metal chains and served as a measure of the rate of grinding. While the mechanism of current generation has not been completely clarified, a parallel study of electrical effects associated with rock drilling indicated that the major driving force is the zeta potential*, while oxidation of the freshly fractured material provides the remainder of the current. Both of these effects are sensitive to the exposed surface area, and it might be expected that the observed currents would provide a measure of the grinding rate. This "surface area" effect allows the measurement system to be used in any type of rock or ore grinding. Some studies with sulfide copper ore indicate higher currents than those observed with silica sand, but the overall pattern of current versus grinding rate is essentially the same. Experimental Studies The experimental system was set up as shown in [Fig. 1]. The initial studies were done with silica sand because of its availability and for comparison with the data of reference 1. The sand was sized to an average of 425 /MI (40 mesh); a typical charge consisted of 2.5 kg (5.5 lb) of sand, 0.85 L (0.22 gal) of tap water and 2.5 kg (5.5 lb) of iron grinding balls 25.5 mm (1 in.) in diameter. For a typical "run," the mill was loaded with a fresh charge of ore and operated for periods ranging from 10-60 min. During this time, the current was measured with a Keithley 417 picoammeter and a chart recorder. At the end of the grinding period, the charge was removed and sieved to recover the 425, 300, and 212 µm (40, 50, and 70 mesh) fractions which were then drained and weighed. A fresh charge was used for each test and every attempt was made to keep the draining and weighing processes uniform.

Jan 1, 1985

By Andrew B. Crichton

The oft repeated statements of the United States Geological Survey and the Bureau of Mines that the coal reserves in the United States are sufficient for 3000 yr have given us all a sense of security hard to disturb. If we discount that statement 66 pct, which to the uninformed seems drastic, we would still have 1000 yr supply, and why should mortal man worry about the difference between 1000 and 3000 yr supply of coal. That is exactly the attitude which many who should be better informed have taken toward the problem. A recent brochure "How Strong is America," by the New York Herald Tribune, states there is "no end to coal * * The truth is staggering * * U. S. has enough coal to last 4000 years." What the statement actually means is that the calculator got as high as 4000 yr and decided it was pointless to go on. The President's ERP Committee on Natural Resources and others in high places continue to assure the country of our superabundance of coal. In the past few days the "Wall Street Journal" carried an item from Director Boyd of the Bureau of Mines saying that within a hundred years we would need to triple our annual coal production of 600,000,000 tons, making it 1,800,000,000 tons annually, and at that rate we had coal reserves sufficient for 1000 yr. In my humble opinion, we cannot now or ever produce 1,800,000,000 tons annually, because we do not have the reserves in the producing districts to do it. Past reckoning in geology may be, as in the Psalms, a thousand years are as yesterday, but the future must be considered more realistically, and " a thousand years" is really a "heck" of a long time. In most developed areas, mining engineers can count the coal reserves in terms of a hundred years or less, and in some important areas now producing large annual tonnages in much under fifty years. They think when they hear of thousands of years coal reserves, reference is made to distant fields and "the fields are always green far away." The estimates of our coal reserves upon which we are relying were made in 1907 under the direction of Marius R. Campbell, Director of the United States Geological Survey. It is apparent this work was not well done. It was necessarily based on rather meager data because of the little development and prospecting as of that date. There were mistakes in the correlation of seams, and they assumed the continuity of thick seams throughout large coal bearing areas where coal often did not exist. The overstatement of coal bearing areas in the Campbell figures are common. The 1907 survey includes all coals over 14 in. in thickness to a depth of 3000 it. The total coal reserves according to Campbell were 3157.2 billion tons. The production and mining losses that year would amount to 570 million tons and

Jan 1, 1949

By C. D. Young

IN an effort to employ cast steel of a stronger structure than that found in the annealed steel castings, the possibilities of heat treatment which will increase the strength without materially decreasing the ductility may be resorted to. The following abstracted report (which is a brief outline of what has already been clone by the Pennsylvania Railroad Co. at its shops and Test Laboratories at Altoona, Pa.) is of material interest, as it indicates what may be done with steel castings when properly treated, thereby permitting in railway service greater strength of cast steel parts without any increase in weight or space. The question of the heat treatment of alloy steel castings is not taken up in this paper. The obscurity formerly surrounding the heat treatment of steel has been for the most part removed by the development of our knowledge of the critical points of steel, pyrometers, furnace construction, and the testing of the finished product. The operations of the heat treatment proper will be taken up under the heads of (1) heating for quenching; (2) quenching; (3) drawing. (1) Heating for Quenching.-Heating for quenching is best conducted slowly, especially in the case of castings of variable thickness. Cracks may occur either in heating or in cooling, clue to different temperatures at different points of the casting. The castings should be thoroughly soaked at the maximum temperature (generally 1,500° to 1,600° F.), 1 hr. being sufficient for sections 1 ft. in thickness. The minimum temperature which will produce the desired hardening effect will, in all cases, be found to be the most satisfactory, as the grain coarsens when the critical range is exceeded to too great an extent. All temperatures should be governed by a checked pyrometer with the hot junction to the heated object, and with several couples in a large furnace to insure a uniform temperature.

Jan 2, 1914

By E. J. Rapperport, J. P. Pemsler

The thermodynamic activities of zinc in six solid solution Cu-Zn alloys ranging from 5 to 35 at. pct Zn were determined experimentally in the temperature range 400° to 600°C. This low temperature investigation was canducted in order to evaluate techniques developed to utilize the inherently high sensitivity of atomic absorption flocesses in the measurement of thermodynamic activities. Analytical expressions ,for the activity and actizlity coeflcient are given for the six alloys in the temperature ranges investigated. RELATIVELY few experimental methods are available for investigation of thermodynamic activities of alloys, especially in the solid state. The techniques most frequently used have been the electrochemical potential and the effusion methods, both of which have severe limitations in many instances. It is therefore desirable to expand the ability to perform such measurements in order to obtain new information as well as to provide an additional independent verification capability. In this work, we present a significant improvement in the spectrophotometric method for sensing small vapor pressures in static absorption cells. Similar techniques have been used previously;1"5 however, applications had been limited to relatively high pressures, often greater than 1 torr. Prior investigators have, for the most part, used broad spectral sources such as xenon or mercury lamps, and high intensity arcs. Hollow cathode sources were first suggested in 1956 6 and were used soon afterwards.4'5 These sources offer significant improvements in sensitivity and freedom from interfering spectral lines.'-' EXPERIMENTAL High purity zinc was obtained from Cominco Products, Inc., and copper from American Smelting and Refining Co. Both elements were of 99.999 pct purity. Copper-zinc alloys were vacuum melted in a high fired carbon crucible with each alloy pulled from the melt as a 4 -in. diam bar. The bars were swaged to -1/4 in. rods and vacuum annealed for 160 hr at 800° + 1°C. Samples for gross chemical analysis were taken at intervals along the length of the rods to ascertain the axial zinc gradient. Electron microprobe analysis of homogenized specimens indicated that the alloys had uniform compositions over their cross sections on a macro (200 p) and micro (1 u) scale to better than *1 pct (20) of the gross composition. This tolerance was determined by counting statistics, rather than assured composition fluctuations. All SiO 2 windows were high-ultraviolet-transmission grade to minimize intensity losses. Silica absorption cells were scrupulously cleaned consecutively in organic solvents, dilute HF, and distilled water before use. The empty cells were then flamed while under a dynamic vacuum, cooled, and removed to an argon-filled glove bag. Alloy pieces were cut and filed in the glove bag to produce fresh surfaces, and then loaded into the cells. The loaded cells were temporarily sealed, removed from the glove bag, reevacuated to 10-5 torr or better, and permanently sealed. The instrument used is schematically shown in Fig. 1. The spectral emission from a commercially made hollow cathode lamp (A) of a selected element is focused through an absorption cell (B) inside a well-controlled furnace (C). The intensity of the transmitted beam is measured using the spectrometer* (D) 'Techtron model AA4 atomic absorption spectrometer. which contains a grating (E) that disperses the light prior to impingement on the photomultiplier (F). The monochromator grating is adjusted so that only the wavelength of interest is measured. The power supply delivered an interrupted voltage to the lamp, causing a chopped radiation output to be transmitted. The detector read only the intermittent component of radiation incident upon it, so that all continuous noise signals (furnace radiation, and so forth) were eliminated. Three recording thermocouples contained in the muffle furnace were positioned along the length of the absorption cell: one at each end and one at the center. An effort was made to keep the ends of the cell several degrees hotter than the center to avoid window condensate. Appropriate thermal corrections were then necessary to relate cell pressure to radiation attenuation. Water-cooled heat shields, as shown in Fig. 1, were found to aid signal stability by protecting the hollow cathode and the photomultiplier from furnace radiation. The furnace had a 2-in. diam muffle, Kan-thal wound, with SiO 2 windows at its ends to minimize convective effects. The hollow cathode radiation was masked and focused to form a conic beam that was a maximum of { in. diam within the furnace. Thus, the 1.5 in. diam absorption cell easily contained the entire beam. The furnace was mounted on ball-bearing slides with positive positioning detents. This arrangement allowed the removal of the entire furnace assembly from the radiation path, position [I], Fig. 1, so that frequent sampling of the unattenuated beam intensity could be obtained. In all cases the beam intensity was kept constant to 0.1 pct as judged by readings taken immediately before and immediately after data collection. Only data for absorptions of less than 80 pct were utilized, as systematic deviations from linearity were found for greater absorptions.

Jan 1, 1970

By William A. Kapp

INTRODUCTION Coal is being mined from beneath residential areas, structures, bodies of water and other surface features in the coalfields to the north, south and west of Sydney. The particular problems faced by mine operators in these areas vary considerably due to differences in the overlying strata, the variation in the depths of cover and also depend on the number of seams being mined. Detailed subsidence work first commenced in the Southern Coalfield in 1965 and is now being carried out over areas of extraction at roost collieries. The analysis of the results of the early investigations and of the work which continues in other areas has shown that there is a consistent relationship between subsidence and mine geometry and has led to a reliable empirical method for the prediction of subsidence. In addition, particular aspects of each of the studies in the Southern Coalfield results in a clearer understanding of strata movements and of the resulting subsidence. The features of a subsidence trough apply generally to all areas but the magnitudes of specific features vary according to the stratigraphy of the particular coalfield. The aim of the subsidence work is to quantify the effects of subsidence for a range of mining geometries and mining conditions to enable the maximum safe recovery of coal from beneath surface features. The importance of local subsidence investigations is becoming more evident to mine operators and to authorities or organisations with surface interests. The subsidence work also provides important information on the stabilities of pillars of coal which remain unmined between panels of extracted coal. These pillars are not extracted either because of poor mining or geological conditions, or because pillar extraction is not part of the particular mining operation. Subsidence studies over these coal pillars clearly establish whether the pillars have remained stable or have failed to support the overlying strata. With subsidence studies continuing over several years, it is possible to assess the stabilities of these pillars on a long term basis. BACKGROUND TO THE STUDY OF SUBSIDENCE Geographical setting Most of the black coal production in Australia comes from the Sydney Basin. The coal seams extend for approximately 350 km along the coast of New South Wales and inland for distances up to 150 km. The City of Sydney is located near the centre of the coastal extent of the Basin where coal has been mined at a depth of 900 m. The Sydney Basin is part of the Main Coal Province of NSW and is divided into several coal- fields. The Southern Coalfield to the south of Sydney contained 15 operating mines and produced 12.7 million tonnes of raw coal during the 12 months to June 1981. The collieries discussed later are shown in Fig. 1. The prominent topographical feature of the area is the Illawarra Escarpment which rises to 400 m above sea level, or 300 m above the coastal strip along the South Pacific Ocean. The escarpment is mainly sand- stone and the weathering of the cliff line has resulted in a covering of talus material at its base. Several collieries are located near the seams which outcrop along the escarpment. The city of Wollongong is located in a scenically attractive area on the coastal plain. The suburbs of Wollongong extend north along the coastline, south to beyond Lake Illawarra and west to the lower slopes of the escarpment. The Illawarra Escarpment forms the eastern boundary of the Woronora Plateau. On a regional scale the surface dips gently to the west and thus forms a watershed for the rivers, most of which flow in a general north westerly direction, sometimes forming steep gorges in the sandstone. These rivers join the Nepean and Hawkesbury River system and flow into the Pacific Ocean north of Sydney. Seven dam have been constructed over the Southern Coalfield (Fig. 1) and with one large dam further to the west, their stored waters provide the needs of the Cities of Sydney and Wollongong and the surrounding districts. A large part of the area affected by mining is the undeveloped bushland of the associated catchment areas. In general no special precautions have been required with respect to subsidence with the exception of the dam structures and stored waters. With the increase in coal mining activities and the expanding residential development south of the City of Campbelltown in the outer Sydney Metropolitan area, subsidence is becoming an increasingly important area of research. Structures which have been affected or considered are townships and extensive residential areas, buildings of historical importance, major tollways, and a high pressure natural gas pipeline. The subsidence effects of mining beneath natural features within national parks is coming under study as mining approaches these areas. Geological setting The coal seams of the Southern Coalfield lie within the Illawarra Coal Measures. They contain high rank coking coal used in the local steel industry and for export. The Bulli Seam is mined extensively through- out the Southern Coalfield with the lower Wongawilli Seam being second in importance with regard to coal production. The top of the Bulli Seam is taken to be the marker horizon between the Permian Coal Measures

Jan 1, 1982

By Heinrich Siemes, Christa Hennig-Michaeli

The effect of temperature and orientation on the plastic behaviour of an impure titaniferous magnetite from Palabora (South Africa) has been investigated. Prismatic specimens were axially compressed along [001], [110] and [111]. The strain rate was about 3.10-6 s-1. Trace analyses of glide lines revealed a plastic anisotropy due to glide on several non-equivalent crystallographic planes. At 200 °C {111}<112> twinning, {111}<110> slip and {100}<011> slip have been operative. At 400 °C {111}<110> slip and {100}<011> slip are primary glide modes. {110}<110> slip seems to be a secondary glide mode in the [001] specimens at 400 °C. {III}<112> twin glide generates isolated sharp lamellae. {111}<110> slip produces smooth surface striae. {100}<011> slip is characterized by coarse and partially wavy slip bands. {I10}<110> slip displays faint remarkably equidistant glide lines. The critical resolved shear stresses (CRSS) decrease with rise of temperature. The CRSS ratio of slip on [100) to slip on {lll} increases from 1.1 at 200 °C to 1.4 at 400 °C giving rise to a distinct flow anisotropy of the magnetite at 400 °C.

Jan 1, 1982

By Nathaniel Arbiter

THE separation of minerals by flotation can be regarded as a rate process, with the extraction of any one mineral determined by its flotation rate, and the grade of concentrate by the relative rates for all the minerals. So regarded, the significant variables for the process are those that control the rates. These variables are of two types, the first describing the ore and its physical and chemical treatment prior to flotation and the second characterizing the separation process in the cells. This paper will examine the variation in rates for a group of separations, will show that a simple rate law appears to govern, and will consider the relation of the control variables to the rates. The use of rate constants for evaluation of performance and efficiency will be discussed. Flotation involves the selective levitation of mineral and its transfer from cell to launder. The flotation rate is the rate of this transfer. It may be defined by the slope of a recovery-time curve for any cell in a bank, or at any time in batch operation. The objective in flotation rate study is an equation expressing the rate in terms of some measurable property of the pulp. This can be either the concentration of floatable mineral in weight per unit volume1,2 or a relative concentration, which will be a function of the recovery." A rate equation for an actual flotation pulp will contain at least two constants, both to be determined from the data. One of these, the initial concentration or proportion of floatable mineral, is not necessarily equal to the feed assay because of nonfloatable oversize or locked particles." The other, a rate constant, is a measure of proportionality between the rate and the pulp property on which the rate depends. The value of the rate constant will be determined by the values of all variables which control the process and will be changed by significant changes in any of them. It is, therefore, a direct measure of performance. Where recovery or grade change continuously with flotation time, the rate constant will be independent of time and will characterize the entire course of the separation. Development of Rate Equations Rate equations can be developed either by analysis of the mechanism of the process or by direct fitting of equations to recovery-time data. Sutherland&apos;s attempt by the first method&apos; suggests that the effect of particle size variation on the rate complicates the derivation of a simple equation applicable to an ore pulp. A further problem with an ore is the concentrate grade requirement, which usually involves a variable rate of froth removal. Thus the final rate for any cell may depend on the froth character and froth height, as well as on the pulp composition. This does not imply that each cell cannot reach a steady state2 in which the rate will depend ultimately on pulp composition. The second method is the fitting of rate equations consistent with the necessary boundary conditions* to experimental recovery-time curves. On the assumption that under constant operating conditions the flotation rate is proportional to the actual or relative concentration of floatable mineral in the pulp, a generalized rate equation may be expressed as follows: Rate = Kcn [I] where K is the rate constant, c is some measure of the quantity of floatable mineral in the pulp at time t, and n is a positive number. In previous rate studies, the value of n has been taken as 1, either by direct assumption," or as a result of the hypothesis that bubble-particle collision is rate determining.&apos; A first order equation results, which after integration in terms of cumulative recovery R, leads to Loge A/A-R = Kt [2] The quantity A is the maximum possible recovery with prolonged time under the conditions used. No conclusive proof for the validity of this equation in flotation has been advanced. The evidence cited in its support consists entirely in the demonstration that it appears to apply to a limited number of recovery-time curves."&apos; It will be shown subsequently that this procedure is not sufficient to establish the order of a flotation rate equation. The possibility that the equation may be of higher order therefore requires examination. If, in particular, the exponent in eq 1 is assumed to be 2, then after integration there results R = A2Kt/1 + AKt [3] with K again a rate constant and A the maximum proportion of recoverable mineral. Eq 3 may be

Jan 1, 1952

In accordance with the provision of the Constitution, notice is here given to all Members, Associates, and Junior Associates, that the dues of the year 1919 will be payable on Jan. 1, 1919, at the office of the Secretary. The clues for Members and Associates are $12 per year and the sum of. $2 additional is asked of those Members and Associates who desire to &apos;have the two volumes bound in the Institute&apos;s standard binding. The clues of Junior Associates are $5 per year.

Jan 12, 1918

By Gustav S. Preller, A. M. Gaudin

For the past 20 years it has been generally accepted that the flotation process is made possible as a result of the action of certain chemical substances on the surface of the mineral particles. In fact, it has been suspected, but never demonstrated, that this action produces on the mineral particles a layer of chemosorbed ions that is of the order of magnitude of a monolayer; that is, of a layer one ion deep. It is the purpose of this paper to present certain initial results on the relationship of mineral surface area and collecting-agent requirements. Up to about eight years ago, all considerations involving the surface area of finely divided solids were limited by the fact that there was not in existence any general method for determining the extent of such surfaces, especially on mixed materials. However, the method invented by Brunauer, Emmett and Teller1 has changed the situation by providing a general procedure (BET method). This procedure has already been applied to mineral dressing problems2-4 and has now received confirmation as to accuracy by an entirely independent method.&apos; This confirmatory or "absolute" method uses a very tedious calorimetric procedure and is not suitable for routine determinations; but still another method6 now is available to relate the results obtained by the BET method to those obtained calorimetrically. This Harkins-Jura method has the advantage that it does not require an assumption concerning the parking area per molecule of adsorbate (nitrogen). In the original work by Brunauer, Emmett and Teller the parking area per nitrogen molecule was calculated to be either 13.8?2 or 16.2?2. according as the nitrogen was assumed to be closely packed as in crystalline nitrogen or loosely packed as in liquid nitrogen. The initial experimental work from this laboratory2 supported the 13.8?2. figure; that work was done on glass spheres of measurable geometric area and it was based on the assumption that the glass spheres were truly smooth; any deviation from this assumption would, of course, have made the calculated parking area higher. Harkins and Jura&apos;s results suggest that the parking area is not the same for all substrata; furthermore, the average of the 87 measurements cited indicates a "preferred" parking area for the nitrogen molecule of about 15.2?2., with extreme values at 13.7 and 16.9?2. The average of 21 determinations made in connection with this study7 gave a parking area of 14.3 ± 0.27?². In view of this result and of the Harkins and Jura

Jan 1, 1947

By Gustav S. Preller, A. M. Gaudin

For the past 20 years it has been generally accepted that the flotation process is made possible as a result of the action of certain chemical substances on the surface of the mineral particles. In fact, it has been suspected, but never demonstrated, that this action produces on the mineral particles a layer of chemosorbed ions that is of the order of magnitude of a monolayer; that is, of a layer one ion deep. It is the purpose of this paper to present certain initial results on the relationship of mineral surface area and collecting-agent requirements. Up to about eight years ago, all considerations involving the surface area of finely divided solids were limited by the fact that there was not in existence any general method for determining the extent of such surfaces, especially on mixed materials. However, the method invented by Brunauer, Emmett and Teller1 has changed the situation by providing a general procedure (BET method). This procedure has already been applied to mineral dressing problems2-4 and has now received confirmation as to accuracy by an entirely independent method.&apos; This confirmatory or "absolute" method uses a very tedious calorimetric procedure and is not suitable for routine determinations; but still another method6 now is available to relate the results obtained by the BET method to those obtained calorimetrically. This Harkins-Jura method has the advantage that it does not require an assumption concerning the parking area per molecule of adsorbate (nitrogen). In the original work by Brunauer, Emmett and Teller the parking area per nitrogen molecule was calculated to be either 13.8?2 or 16.2?2. according as the nitrogen was assumed to be closely packed as in crystalline nitrogen or loosely packed as in liquid nitrogen. The initial experimental work from this laboratory2 supported the 13.8?2. figure; that work was done on glass spheres of measurable geometric area and it was based on the assumption that the glass spheres were truly smooth; any deviation from this assumption would, of course, have made the calculated parking area higher. Harkins and Jura&apos;s results suggest that the parking area is not the same for all substrata; furthermore, the average of the 87 measurements cited indicates a "preferred" parking area for the nitrogen molecule of about 15.2?2., with extreme values at 13.7 and 16.9?2. The average of 21 determinations made in connection with this study7 gave a parking area of 14.3 ± 0.27?². In view of this result and of the Harkins and Jura

Jan 1, 1947

By Arno C. Fieldner

PROBABLY no large industry in the United States is in greater need of technologic research leading to economic improvement than the coal industry. It has suffered severely from increasing substitution of petroleum and natural gas for coal and its products in the fuel markets. Part of this substitution is due to the greater convenience and suitability of liquid and gaseous fuels for special purposes, and part of it has been due to the extensive research programs conducted or sponsored by the petroleum and allied industries in the past 2 decades. A recent estimate1 places these expenditures at $100,000,000 a year.

Jan 7, 1951

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Jan 1, 1938

By W. E. D. Stokes

WHEN history is written, the year of the blitzkrieg will go down as giving aviation its greatest impetus. No perceptible drop in military business, even with cessation of hostilities abroad, seems likely. The combined air force of the U. S. Army and Navy on July 1 was approximately 5000 planes, as against 50,000 planes currently discussed as necessary. Observers at the fronts are giving designers and metallurgists back home some realistic ideas to shoot at. As conventional runways reveal the location of airports to enemy raiders, operating bases are now masked. Difficult, often muddy and rough landing strips, requiring strong landing gears and exacting take-off and maintenance characteristics, are the specifications of war. This is "old stuff" to numerous mining operators in mountainous jungle areas where gradients up to 8 per cent are in more or less common use in such mountain landing fields. Although airplane building is not readily adaptable to assembly-line methods, a trend towards mass productions is now in evidence. Incidentally, a U.S. Army "Curtiss P-40" requires about 10,000 man hours of work to complete and is tied together with 75,000 rivets. During the past year a number of aircraft factories adopted the "shot-

Jan 1, 1941

By Charles L. Harness

MINERAL pigments give color, opacity, or body to paint, stucco, plaster, mortar, cement, linoleum, rubber, and similar materials. They must be finely divided, substantially insoluble, and generally inert for such uses. Inorganic pigments may be grouped somewhat arbitrarily as follows: A. Natural mineral pigments dug from the earth. 1. Those consisting mostly of iron oxides, such as hematite and Limonite. 2. Those containing large amounts of clay or noncoloring matter, such as ocher, sienna, umber, and colored shales. 3. Those whose color is not due to iron oxide, such as Vandyke brown (including sap brown, an alkali extract of Vandyke brown), graphite, and terre verte. B. Synthetic pigments. 1. Those containing iron oxide as the essential coloring matter, such as "pure" reds, yellows, and blacks, and Venetian reds. 2. Those not containing iron oxide, such as lithopone, zinc white, titanium white, ultramarine, Prussian blue, and the like. Only groups A and B-1 are discussed here. All the pigments in those two categories, except Vandyke brown, sap brown, and some of the carbonaceous slates, contain, as their essential color constituents, one or more compounds of iron; namely, an oxide-Limonite (2Fe203.H20 to Fe203.4H20), hematite (Fe203), or magnetite (Fe203.FeO). Man&apos;s first permanent coloring materials were earth pigments- the ochers, iron oxides, and various iron-stained clays. The caves occupied by prehistoric man and the arts of the ancient Chaldean, Egyptian, and Cretan civilizations bear mute witness to this and also testify to the beauty and permanence of earth colors. As the Roman era developed, earth pigments were pushed from their pre-eminent position by white lead and red lead and in the Middle Ages by zinc oxide. Today, of course, the number of acceptable pigments is very large and on a value basis the group of natural and synthetic earth pigments stands

Jan 1, 1949

By E. M. Wise

IT is known qualitatively that the recrystallization behavior of platinum is dependent upon the amount and particularly the nature of impurities present, the amount of prior cold-work and the annealing time and temperature, but published data are limited. A study of the effects of cold-work and annealing temperature on the hardness of a sample of very pure platinum was undertaken to supply part of this information. CONCLUSIONS 1. The platinum used in the experiments had a probable platinum content of 99.99+ per cent, and had the following mechanical properties in the completely softened condition: Vickers hardness number, 39.4; ultimate tensile strength, 19,340 lb. per sq. in.; proportional limit, very low; elongation, 37.5 per cent in 2 inches. 2. The effect of cold-rolling on the hardness is shown in Fig. 1. The hardness increases rapidly up to 10 to 15 per cent reduction and more slowly up to about 75 per cent reduction, beyond which it increases more rapidly. 3. The recrystallization temperatures for a 15-min. anneal, after cold-rolling 19, 39.5, 50.8, 66, 80.5 and 89.5 per cent reduction in thickness, are approximately 765°, 700°, 635°, 600°, 545° and 425°C., respectively. 4. The recrystallized grain size of this platinum is a function of the percentage of cold reduction, being larger the lower the percentage of cold-work. 5. Little or no grain growth resulted from subjecting recrystallized samples to higher temperatures or annealing for longer times.

Jan 1, 1940

By F. B. Foley

With the possible exception of high-speed tool steel, the service demanded of rock-drill steel is probably more precarious than that of any other tool steel. Unaided by the helpful influence of alloys and dependent solely on the efficiency of the heat treatment applied for its ability to stand up under the abuse to which it is necessarily subjected, rock-drill steel is worthy of intensive study. For its heat treatment the smith has his fire, an oil fire perhaps, and a quenching bath; he judges the temperature by his eye. There are shops where excellent drills are produced by this method, but there are more where the product varies between very good and very bad. Pyrometers with specially designed equipment are used in but few cases. But the responsibility for rock-drill failures does not rest entirely on the smith. Little has been published concerning the efficiency of drill steel, though the Canadian Mining Journal1 published the breakage record of a large number of hollow drill steels. This record indicates that most of the breaks occur during the early life of the steel and that as a given batch of steel is used, its percentage of breakage decreases. This fact, the article points out, is at variance with the much-talked-of theory concerning failure from crystallization and fatigue. It seems to point to the fact that the greater part of drill-steel failure from breakage is traceable to faulty manufacture, resulting in defects in the bore of the hollow steel from inclusions of slag or oxides, etc. As this defective material meets with early failure, it is weeded out. In addition, some breakage may be caused by a lack of proper condition with respect to heat treatment of the steel as furnished by the manufacturer. In the article mentioned, the breakage of 1/14-in (3.17 cm.) hollow steel is given as 4.62 per cent. for the early period of use, with a decrease to 3.35 per cent. and 2.67 per cent. with time in use. For 1-in. (2.5 cm.) hexagonal hollow steel, the figures for breakage are 0.18 per cent. and 0.06 per cent.; attention is called to the great decrease in breakage by the use of smaller diameter drills. The figures are given as so many drills sharpened, but there is no way of determining the percentage of

Jan 1, 1922

H. McD. Lorain says in regard to his son, Captain S. H. Lorain; "It might be of interest to you to know that, although but 23 years of age, my son is captain of Co. C 107th U. S. Engineers, Thirty-second Division. He landed in France about the middle of February, 1918, and was commissioned a captain on Sept. 5, 1918. He started for the front the latter part of June and was there continuously until the armi-stice was signed, on Nov. 11, being on the front lines in the Argonne at that time. He is, at present, in Coblenz, Germany, with the Army of Occupation. Although much of his work was clone under artillery and machine-blur fire, as well as airplane bombs and fire, he escaped without a scratch. Once, he was the only one of a group who escaped without a scratch from an exploding shell." Lieut. Bernard H. Lasky, on a post card received a short time ago, gave the following brief record of his military career. "I have been serving with the 316th Engineers in France and in Flanders and had the satisfaction of being in the last big drive in Belgium which sent the enemy beyond Avaenarde and did much to increase their gait on their way to Brussels. On Oct. 1, I received a commission as second lieutenant and was assigned to Co. C of the 316th Engineers."

Jan 3, 1919

By H. A. Bourne, E. L. Haden, D. R. Reichmuth

A circumferential-toothed bit with novel tooth form gave improved penetration performance. In this design the exterior flank of all teeth were vertical when in rolling contact with the hole bottom. Rock chips were generated by the interior flank of the tooth displacing the rock inwardly and downslope toward the center of the hole. A unique two-cone laboratory bit assembly enabled evaluation of numerous cone and tooth configurations. Some of the variables investigated, in addition to weight on bit, rotary speed and rock type, were tooth interference, percent tooth, hole bottom angle, attack angle and relief angle. Most tests were conducted dry on a brittle synthetic sandsone or a ductile quarried limestone. Tooth configurations were found to be more significant in the ductile material. This was attributed to the deeper tooth penetration before rock failure. These studies showed that the attack angle (angle beween interior flank of the tooth and rock surface) was the controlling variable; changing the tooth configuration from the assymetric or semi-wedge to the more conventional symmetric or wedge form reduced penetration performance; and penetration performance of circumferential-type cutters was directly proportional to rotary speeds up to 200 rpm. INTRODUCTION Much of the published literature on rock-chisel interactions describe experiments wherein symmetrical wedges are vertically loaded or impacted against a smooth rock surface.1-6 are is usually taken to insure that the indentation is not made near the edge of the rock specimen less erroneous data be obtained. The literature describes relatively few studies in which the investigator deliberately attempted to take advantage of an edge or free surface. In contrast, anyone who chips ice or breaks up a concrete sidewalk almost always works near an edge. Chisel "indexing," which has been considered by some investigator1,2,6,7 makes limited application of an edge or free surface. Probably the best documented investigation into applying this idea to drilling was that of Drilling Research Inc. at Battelle Memorial Institute.&apos; Their "annular wing" percussion bit consisted of paired asymmetric chisels oriented so as to produce and move chips to the center of the hole. They predicted that the lowest energy requirement for chip generation would be achieved with a stepped hole bottom having a median angle of 45" to the horizontal. Results from limited tests showed that approximately 50 percent of the rock fragments were large and semi-circular in shape, as would be expected by a chisel impact near an edge. The remaining 50 percent were fine chips produced by the chisels in re-establishing the steps or ledges. Initial penetration rates with this bit were high, but they rapidly decreased. This was the result of excessive tooth wear caused by the constant friction on the gauge surfaces. The basic idea — circumferentially placed asymmetric chisels — still appears to have merit. If the concept could be applied to a rolling cutter bit, two of the shortcomings of the fixed chisel design could be overcome: (1) reduction in tooth friction, and (2) greatly increased cutter surface. Adapting asymmetric chisels to cutters rolling on an inclined hole bottom is restricted by bit geometry. The basic elements of roller rock-bit construction prevents the practical attainment of a 45" hole bottom angle. Nonetheless, experimentally it was considered desirable to investigate the influence of hole bottom angle to at least 40". This paper describes the laboratory studies conducted in evaluating the circumferential-toothed roller cutter rock bit. EXPERIMENTAL APPARATUS AND PROCEDURE BIT ASSEMBLY The cost of constructing a sufficient number of conventional three-cone rock bits to investigate circumferential cutter performance was prohibitive. Instead, a novel two-cone laboratory assembly which used an external bearing system was designed and constructed. The external bearings made it possible to alter the journal bearing angles and thus allow a wide flexibility in cutter configuration. Fig. 1 shows the laboratory bit assembly, the various bearing mount plates and the appropriate roller cutters for drilling shallow holes having hole bottom angles of 0, 10, 20, 30 or 40". The bit was limited to a drilling depth of 1 1/2 in. at the gauge teeth and a hole diameter of 43/4 in. This more or less intermediate size bit was chosen because it gave a more realistic match between bit teeth and the rock than would a microbit. Also, the rock sample size required was convenient and easy to obtain. CIRCUMFERENTIAL CUTTERS The tooth configuration used in our initial studies is shown in the upper half of Fig. 2. All cutters used in this series had the same tooth form — 43" included tooth angle, 2" positive relief angle and a horizontal tooth flat width of 1/32 in. Each cone cuts alternate rows except for the gauge row. The row-to-row spacing in view was 1/4 in. Static loading tests conducted earlier with asymmetrical chisels had been used to establish this spacing. These tests showed energy requirements for chip production increasing rapidly as the distances to the edge increased beyond

By R. T. Keyes, R. B. Clay, L. L. Udy, V. O. Cook, M. A. Cook

Based on compressibility and stress wave velocity in rock, initial explosive loading conditions, the thermochemistry of the explosive and reasonable description of the pressure-distance relations behind the shock front, the distribution of energy between the products of detonation and the burden are estimated as a function of time for various loading conditions. These include &apos;powder factor&apos;, loading density A (fraction of borehole occupied by explosive), and the explosive itself. Factors responsible for rock fragmentation are discussed in terms of: 1) &apos;release wave fracturing&apos;, 2) &apos;shear wave fracturing&apos;, and 3) &apos;release of loading fracturing&apos; or &apos;rock bursting&apos;. Release wave fracturing appears to occur only in the immediate vicinity of free faces, and shear wave fracturing only adjacent to the borehole under normal blasting conditions. Rock bursting is thus considered to be the most important means of rock fracturing in blasting. Means for maximizing it are considered. The most important factors to consider include maximum available energy A, explosive density p1, loading density A and &apos;powder factor&apos;. Maximum efficiency is attained in general by maximizing the A . p1 product. This is achieved by using high density explosives at a loading density of unity (A = 1.0). Ways for achieving this condition are discussed. The bulk-handled aluminized slurry blasting agents have the desired properties for achieving optimum conditions for high blasting efficiency based on the theory outlined herein. Factors considered most important in the blasting of rock are: 1) The maximum available energy A, determined by the heat of explosion Q and the mechanical efficiency ?, a factor intimately associated with the mode of application (A = Q at highest gas concentrations). 2) The &apos;borehole pressure&apos;, pb, or the maximum pressure developed in the borehole after passage of the detonation wave and before the burden has had time to move or become compressed appreciably. (Owing to the short duration of the detonation wave at any particular point in the borehole, the fact that the explosive may not always fill the borehole completely and the further fact that the burden may not actually see the detonation pressure, the borehole pressure is considered more significant than the detonation pressure p2 as a performance factor in borehole blasting.) The borehole pressure is determined by the explosion or adiabatic pressure, p3 , and the loading density, A, or the fraction of the borehole filled by explosive. 3) The physical conditions important in the application of the explosive are: a) The &apos;powder factor,&apos; (We/Wr), or the ratio of the wt of the explosive to that of the rock being blasted expressed as lb per ton or more generally in lb/cu yd. b) The bulk density, p1 ?. c) The &apos;burden&apos; or &apos;line of least resistance,&apos; the spacing between boreholes, the geometry of the borehole pattern and sequence of firing. d) The physical and chemical properties of the rock, most significant of which are possible heterogeneties, such as faulting, prefracture, and greater than micro-scale chemical heterogeneities. (This factor is not considered here but deserves a great deal of careful consideration.) All of these factors need to be carefully considered in the most economical engineering of a blast. Here is considered firstly an outline of the present status of dynamic rock mechanics, particularly as it pertains to blasting. The factors pertaining to the most efficient application of blasting agents are also considered, followed by a discussion of methods of application to achieve optimum explosives performance. ROCK MECHANICS Dynamic rock mechanics is currently a rapidly developing science contributing greatly to a better understanding and consequently the more effective application of explosives in blasting of rock.1-22 Basic to the development of the science of dynamic rock mechanics were the advances of Goransen23 and the Los Alamos and NOL groups24-28 concerning rock and stress wave phenomena and the transmission and reflection characteristics of stress waves at interfaces between different media. Basing considerations on this new knowledge as well as new experimental methods of study (ultra-high speed streak and framing cameras and electronic timers) the theory of fracture and failure of solids under impulsive loading by stress waves developed rapidly.14,16,27-29 Also the

Jan 1, 1967

The following is an incomplete list of members and guests who. called at Institute headquarters during the period Dec. 10, 1917 to Jan. 10, , 1918: A. A. Arluck, Camp Merritt, N. J. W. Spencer Hutchinson, Boston, Mass. C. L. Bruns, Santiago de Cuba. William F. Jahn, New York, N. Y. Henry C. Carr, New York, N. Y. C. H. James, Tucson, Ariz. L. J. Charles, Santa Fe, N. M. Charles Janin, San Francisco, Cal. W. A. Deane, Waxhaw, N. C. Joseph T. Kupferstein, Shreveport, Pa. P. St. John Dixon, Rio de Janeiro, Brazil. Richard Lamb, New York, N. Y. M. C. Drake, Marquette, Mich. Walter R. Malm, San Francisco, Cal. F. W. Draper, Petrograd, Russia. W. H. Nash, Buffalo, N. Y. Paul W. Gaebelein, Boise, Idaho. Edmund Newton, Minneapolis, Minn. H. F. E. Gamm, Rutherford, N. J. Guy C. Riddell, Port Pirie, South Australia. C. A. Grabill, El Paso, Tex. Williams A. Simkins, San Francisco, Cal. William H. Hampton, New York, N. Y. Henry B. Smith, Northampton, Mass. R. M. Henderson, Brackenridge, Colo. H. E. Smith, Chicago, Ill. Geo. B. Holderer, Northpines, Ont., H: C. Spaulding, Washington, D. C. Canada. Y. Wakabayashi, Tokyo, Japan. H. H. Howry, Camp Upton. N. Y. A. P. Watt, Mine La Motte, Mo. L. Douglas Anderson", superintendent of the Midvale, Utah, smeltery of the United States Smelting Co., has been appointed manager of smelteries for the company at Salt Lake City, Utah.

Jan 2, 1918