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The Bougainville ore body was discovered in 1964 by a partnership of Conzinc Rio Tinto of Australia and New Broken Hill Consolidated Limited. The ore body is a porphyry copper in a fine-grained quartz diorite, which intrudes andesite volcanics and is located at an elevation of about 2, 000 ft in the Crown Prince Range of Bougainville in the Solomon Islands. Mineralization is predominantly chalcopyrite with some bornite and some gold and silver present. Rainfall in the area amounts to 200 in. (5. 08 m) per year. An active volcano is situated only 24 miles (38.4 km) from the mine, and all structures were designed to resist seismic shocks. An interesting and unique feature of the Bougainville operation was that the overburden was stripped with hydraulic monitors. Bougainville Copper, at 90,000 tpd, is the largest copper concentrator to be designed and constructed at one time; and besides the twin 54 in. by 74 in. Allis-Chalmers gyratories, it features seven 7 ft (2. 13 m) standard and fourteen 7 ft shorthead crushers, in addition to nine 18 ft diameter by 21 ft long (5. 5 m by 6. 4 m), 4250 hp, single -stage ball mills. During the design stage of the Bougainville primary crushing plant, several alternates were studied because the ore was expected to break with the production of large amounts of small-sized particles but with a minimum of fines. These alternates included two 54 in. by 74 in. gyratories with grizzlies, two 60 in. by 89 in. gyratories, and one 72 in. by 96 in. gyratory, the latter two installations without grizzlies. The two 54 in. units with grizzlies were finally selected as giving good reliability and the ability to handle the large amounts of small material in the feed. As originally installed, grizzly bars were set at 8 in. (203 mm); but, in an effort to perform more interparticle crushing in the primary and less rock-on-steel crushing in the secondaries and tertiaries, a test is underway (1976) to determine the effect of using 6 in. (152 mm) grizzly spacings (see write -up on Anaconda's crusher at Butte). The Bougainville ore is mined with 17 cu yd (13 cu m) shovels and hauled in 105, 170, and 210 st trucks to the crushers, which are equipped for two-side dumping. Up to 40% of the ore passes the grizzlies, which are equipped with splitters; and additional horizontal splitter bars are located directly above the spider caps. Because of the grizzlies, the crushers are supported by unique U-shaped enclosures, which are in turn supported by the end and center walls of the installation.

Jan 1, 1978

By P. D. Hunt, I. Johnson, M. G. Chasanov, H. M. Feder

The solubilities of the transition metals from scundium to nickel, inclusive, in liquid cadmium were determined by sampling saturated solutions. At 400°C these solubilities (ppm) are:Sc, Co, 22; Ni, 12000. Relative partial molal enthalpies and excess partial molal entropies at infinite dilution for the solutes Cr, Mn, Fe, and Co are compared zoitlz similar data for the post-transition metals. Three new intermetallic phases, ScCd,, TiCd, and Ti,Cd, encozaztered in the course of this work, are reported. Measurements of the solubility of the 3-d transition metals in liquid cadmium have been reported' only for iron. The present measurements for the entire series were made as an aid to the study of the systematics of alloy formation. EXPERIMENTAL The apparatus and sampling techniques used for the solubility determinations were similar to those described by Martin etal.' Temperatures were measured with a Pt/Pt-10 pct Rh thermocouple sheathed in recrystallized alumina and immersed in the solution. The thermocouple was recalibrated periodically during the course of the investigation. Thermocouple voltages were measured to the neares 0.01 mv with a Rubicon model 2732 potentiometer. A scaled-down version of the apparratus, in which the total metal charge was about 90 g, was employed to measure the solubility of scandium. Two-mm ID Vycor filter pipets were used to sample the scandium-cadmiurn solutions; the samples weighed about 2g- Cadmium with a minimum purity of 99.95 pct was used; the only impurities detected spectrochem-ically were C'u and Mg, both less than 10 ppm. The purities of the other metals were: scandium, 99.9 pct; titanium. 99.92 pct; vanadium, 99.9 pct; chromium, 99).99 pct; manganese, 99.9+ pct; iron, 99.90 pct; cobalt, 99.95 pct; and nickel, 99.96 pct. Solidified melts were routinely analyzed spectro-graphically. No significant contamination by the crucible, sampling pipets, or other materials of construction was observed. The entire specimen obtained by sampling was dissolved to preclude analytical errors due to segregation on cooling. Colorimetric methods3 were used for the analyses of cobalt (2-nitroso-l-naphthol), manganese (permanganate), nickel (dimethylglyoxime), titanium (peroxide), and vanadium (thiocyanate).4 Chromium, iron, and scandium were neutron-irradiated to produce The dissolved samples were gamma-counted to determine solute concentration.For the separation of intermediate phases from cadmium-rich alloys selective leaching by saturated ammonium nitrate solutions was employed. The concentration of cadmium in the successive leaches was followed polarographically and the process was terminated when the rate of cadmium dissolution decreased markedly. For the further investigation of intermediate phases 1/4- in. diam powder compacts were pressed at 80,000 psi and annealed in argon at 315" to 330°C for 10 days. Larger quantities of alloys were prepared in sealed tantalum capsules heated in a rocking furnace.

Jan 1, 1962

J. E. JOHNSON, JR., New York, N. Y. (communication to the Secretary*).-The discussion of Mr. Henrich's paper took place at a meeting from which I had the misfortune to be absent, and has just come to my attention. I do not wish to let the subject pass without contributing to it a little information which has come into my possession. In 1912, I published a brief note in the Transactions on The Effect of Alumina in Blast-Furnace Slags, giving in tabular form three typical slags used in iron blast-furnace practice, one of them the high alumina slag of, which Professor Richards speaks. That table is reproduced herewith. TABLE 1.-Composition of Slays. 1 3 3 4 5 6 7 CaO by Neutral Ratio Ratio Ratio Al2z03, SiO2, Differ- Substan- Cat Ca0 a A1203 Ratio Per Per ence, ces, CaO Cent. Cent. Per Per Al2O3 + Sioz Sit: SiO2 Cecil_ Cent. Virginia practice 6.5 36.0 54.0 3.5 1.27 1.6S 1.50 Lake-ore practice 13.5 33.0 50.0 3.5 1.08 1.92 1.51 Special experiment... 36.0 24.7 36.8 3.5 0.59 2.90 1.44 It supplies what seems to me convincing evidence that, for the purposes of the iron blast furnace, alumina can be regarded only as a neutral substance, dissolving in the slag and materially altering its viscosity, but not affecting its chemical activity. In subsequent experience with charcoal furnaces, I found the same rule to hold; that the chemical effect of alumina could be disregarded, providing the ratio of the lime and magnesia to the silica were kept constant. This is exactly the experience which Messrs. Dwight and Mathewson had in copper- and' lead-smelting furnaces. I have never seen any evidence or argument adequate to support the contention of Professor Richards-that the alumina acted as a base in the charcoal iron furnace and as an acid in the coke furnace, and that the function of alumina in iron-furnace slags is inherently different from that in copper- and lead-furnace slags-and I am compelled to differ absolutely with his conclusions in that respect.

Jan 9, 1917

By Alistair Frame

INTRODUCTION In April 1988 discussions were about to start with BP on the possibility of buying their- mineral interests, and it took a long time to arrange a deal. I will talk about RTZ without BP Minerals, look at the new RTZ with BP Minerals, discuss the outlook for metal prices, and review the problems in South Africa and develop the case against sanctions on that country. RTZAT THE END OF 1988 Late in 1987, after completing a worldwide review of our strategic thrust, we decided: 1. That the business focus would be to remain a major diversified natural resource group with interests in related industries. 2. To divest non-core and low growth businesses. 3. To develop a logical acquisition plan. 4. To emphasise organic growth and, most importantly, to enhance shareholders' wealth by capital increase and a progressive dividend policy. Subsequently, we concluded that we should sell our oil and gas business and dispose of several assets for which we saw little growth potential. The disposal programme raised some U.S.$ 1,600 million, representing a significant capital gain on our various investments. All transactions were carried out by our small Head Office staff in London, without the help of merchant banks and lawyers, and we were able to safeguard the employment of our staff in all these operations. The most interesting disposal was of our oil and gas interests. Before the 1987 strategic review, we had ambitions to develop our oil and gas interests and to become the leader in the second division of European oil companies. Our interests, which were purely in exploration and production of oil and gas, consisted of the 100% owned RTZ Oil and Gas company, and 29% of LASMO, a publicly quoted U.K. oil and gas company with extensive worldwide interests. LASMO, in turn, owned 25% of Enterprise Oil, one of the more successful of the North Sea oil exploration and production companies. Prior to the end of 1987, our intention was to acquire control of both LASMO and Enterprise. However, when we evaluated the net asset value of these two companies based on our oil price assumptions, which were pessimistic for the next five years or so, we concluded that they were both trading in the market at significant premiums. This, taken with the additional premium required for acquisition and the high cost of exploration, led us to conclude that, since funds could undoubtedly be used to better effect, we should sell our interests. Turning to the other major achievements in 1988, we commissioned, and built up to design levels in a very short period of time, the Morro do Ouro gold mine in Brazil. This mine reached its design output of 100,000 ounces of gold per year within three months of start-up, which was on time and under budget. We are now considering a possible expansion of this property. In South America, Utah International, Mitsubishi and RTZ arranged non-recourse finance, with post completion guarantees, for the giant Escondida Project in Northern Chile. Financing came from a group of banks in Japan, Germany and Finland -- the countries to which most of the concentrates will be sold -- and from the LEC. The average output of the mine for the

Jan 1, 1990

By H. M. Chance

Jan 1, 1887

Jan 1, 1923

Jan 1, 1923

By F. W. Osterwald, C. R. Dunrud

Coal mine bumps are a serious hazard to life and property in the mines of east-central Utah. Research into geologic factors associated with these bumps indicates that the bumps are spatially and genetically related to structural and stratigraphic features. Some bumps are directly related to stress accumulation along faults, either from natural causes or as a result of mining. Frictional properties of bedding planes between coal and roof rock, and between some rock units within the roof, directly affect the deformation of coal ribs, and hence the incidence of bumps. These frictional properties are related to the lithology of the roof rocks as well as to the sedimentary structures within the rocks. Commonly a sequence consisting of 1/3 ft of carbonaceous silt-stone, about 2/3 ft of rider coal, and several ft of massive siltstone or sandstone overlies the main seam. The sequence causes difficult roof conditions in the mines, difficult mining conditions and, directly or indirectly, many bumps. The U. S. Geological Survey has studied the relationships between geological features and coal mine bumps at Sunnyside, Utah, for about six years. The work, which was originally undertaken at the request of and in cooperation with the U.S. Bureau of

Jan 1, 1965

No. 526.-At liberty about Mar. 1, 1919. Just returned from France, a Captain of Engineers. Member A. I. M. E., A. I. E. E., 35 years old, technical education. Last six years of civil life as electrical engineer, master mechanic and superintendent of construction with large mining company. Salary expected, $3600 per year. No. 535.-Member, 'mining engineer, technical graduate, age 34 years, experience as engineer for coal companies. Three and one-half years with one company and five years with another. Have good recommendations. Recently honorably discharged, First Lieutenant, U. S. Army. Desires position, preferably in West. Salary last received, $175 per month. No. 555.-Metallurgical engineer having 18 years' experience in positions from chemist to manager, specializing in the smelting of copper, gold and silver ores, and now returning to the U. S. Available about May 15. No. 556.-Manager or superintendent, member, technical graduate, age 44, married, desires position. Good references, 18 years' experience in positions of responsibility and administration. Familiar with development, mining, milling, operation, construction and management. Available on short notice. , No. 557.-Member, technical education, married, age 30. Five. years' experience in construction work, four years superintendent of mining and milling operations. Looking for wider experience. Free Apr. 1.

Jan 4, 1919

By John Batchelder, David W. Blake, Ted. G. Theodore

The Copper Canyon CIL-Au-Ag deposits include two geologically distinct hypogene ore bodies developed primarily in late Paleozoic rocks adjacent to a 38-m.y.old potassic-altered granodiorite, whose emplacement seems to have been controlled largely by the Virgin fault, a north-striking normal fault. Widespread but subeconomic concentrations of sulfides as fracture and cavity fillings in skarn apparently preceded the final intrusion of the granodiorite. In the west ore body, metals reached ore-grade concentrations through the massive replacement of andradite by chalcopyrite, pyrrhotite, and pyrite mineral assemblages. The east ore body probably formed after the granodiorite reached its present level. Data suggest that fluids associated with potassic alteration here were most likely composed of magmatic water mixed with some meteoric water.

Jan 1, 1977

By C. J. Ramsburg

The construction and operation of by product coke-oven plants in America are essential to strong national defense and of the greatest importance to many widely diversified undertakings as well as to safety and health. Carbonization Process Coal used under a boiler or in a locomotive is consumed by combustion. Coal used in a by-product oven is distilled or carbonized, meaning that heat is applied to the coal, out of contact with air, causing a distillation that removes the volatile matter in the coal in the form of gas and vapor, leaving coke as a final product. From the gas and vapor distilled from the coal, many valuable substances can be secured, known as by-products. Hence the name "by-product oven." The simplest example of this principle can be shown by means of an ordinary clay pipe. If powdered coal is placed in the bowl of the pipe, the opening of the bowl is plugged with clay, and the bowl placed in a flame, the bowl will become hot, transferring the heat through the wall of the bowl to the coal within. When this coal attains a temperature of about 750°F., gas will begin to pour out of the pipe stem, and this gas may be lighted. A black liquid also will exude from the stem of the pipe, This is tar mixed with certain other products. What remains in the bowl after distillation is coke. A coke oven corresponds to the bowl of the pipe. A modern oven of the latest type is a refractory brick chamber approximately 42 ft. long and 14 ft. high, averaging 18 in. in width. The oven is filled with coal through closable charging holes in the oven top. Each end of the oven consists entirely of a self-sealing door, the removal of which permits the finished coke to be discharged by means of a power-operated pusher machine. The ovens are built in batteries, and the heating is effected by gas burned in vertical flues in the oven walls, each set of flues heating the walls of two adjacent ovens. Such ovens carbonize a charge of approximately 18 tons of coal in 18 hr., or have a daily throughput of 24 tons per oven, producing about 17 tons of coke. In times of need, the output may be increased by raising the temperature of flues, thus shortening the coking time. There is one great essential difference in principle between distillation in a clay pipe and in a coke oven. In the 18-in. oven, the distillation proceeds from two parallel walls held at a high temperature (over zooo°F.) and instead of all the coal being heated at once, the distillation is a progressive one creeping inward. The average travel of this heating is 35 in. per hour from each vertical wall, and as the volatile products are formed progressively they must push away from the interior toward the wall through the previously formed hot coke, then up the walls and Over the top of the coal charge, finally escaping through the outlet pipe. This results in a cracking of the components into many different and valuable substances not formed when the carbonizing temperature is low.

Jan 1, 1942

By C. J. Ramsburg

The construction and operation of by product coke-oven plants in America are essential to strong national defense and of the greatest importance to many widely diversified undertakings as well as to safety and health. Carbonization Process Coal used under a boiler or in a locomotive is consumed by combustion. Coal used in a by-product oven is distilled or carbonized, meaning that heat is applied to the coal, out of contact with air, causing a distillation that removes the volatile matter in the coal in the form of gas and vapor, leaving coke as a final product. From the gas and vapor distilled from the coal, many valuable substances can be secured, known as by-products. Hence the name "by-product oven." The simplest example of this principle can be shown by means of an ordinary clay pipe. If powdered coal is placed in the bowl of the pipe, the opening of the bowl is plugged with clay, and the bowl placed in a flame, the bowl will become hot, transferring the heat through the wall of the bowl to the coal within. When this coal attains a temperature of about 750°F., gas will begin to pour out of the pipe stem, and this gas may be lighted. A black liquid also will exude from the stem of the pipe, This is tar mixed with certain other products. What remains in the bowl after distillation is coke. A coke oven corresponds to the bowl of the pipe. A modern oven of the latest type is a refractory brick chamber approximately 42 ft. long and 14 ft. high, averaging 18 in. in width. The oven is filled with coal through closable charging holes in the oven top. Each end of the oven consists entirely of a self-sealing door, the removal of which permits the finished coke to be discharged by means of a power-operated pusher machine. The ovens are built in batteries, and the heating is effected by gas burned in vertical flues in the oven walls, each set of flues heating the walls of two adjacent ovens. Such ovens carbonize a charge of approximately 18 tons of coal in 18 hr., or have a daily throughput of 24 tons per oven, producing about 17 tons of coke. In times of need, the output may be increased by raising the temperature of flues, thus shortening the coking time. There is one great essential difference in principle between distillation in a clay pipe and in a coke oven. In the 18-in. oven, the distillation proceeds from two parallel walls held at a high temperature (over zooo°F.) and instead of all the coal being heated at once, the distillation is a progressive one creeping inward. The average travel of this heating is 35 in. per hour from each vertical wall, and as the volatile products are formed progressively they must push away from the interior toward the wall through the previously formed hot coke, then up the walls and Over the top of the coal charge, finally escaping through the outlet pipe. This results in a cracking of the components into many different and valuable substances not formed when the carbonizing temperature is low.

Jan 1, 1942

By John T. Thorndyke

MOST important of the naturalcalcium silicates is the meta¬silicate, CaSi03, known as wollastonite, after W. H. Wollaston. A large deposit of this mineral was dis¬covered some seven years ago near Code Siding, Rademacher mining dis¬trict, Kern County, Calif., the deposit being some 7500 ft. long and of an average width of 300 ft. It consists of a series of small domes striking north¬erly and southerly and with a dip of 81 ° to the south. This deposit was opened up by a series of small quarries showing a potential tonnage of 12,000,¬000 tons, allowing of cheap mining and transportation, the Southern Pa¬cific R.R. spur track being one mile north of the workings. Numerous sam¬ples taken from different points on the deposit, tested by Mr. Brinker, of the California Testing Laboratories, Los Angeles, showed an average as fol¬lows: Per Cent Silica (SiO2) 47.12 Lime (CaO) 41.72 Magnesia (MgO) 2.72 Alkalies (R,0) Trace Iron (Fe.,O,) 1.60 whereas the theoretical composition of of wollastonite, as given by William Shirley Bayley, is as follows: Silica 51.75 Lime 45.25 The tonnage and purity of this de¬posit led to investigations into its com¬mercial development. Early work es¬tablished the fact that wollastonite was ideal for the manufacture of a tough, almost unbreakable, glass similar to Pyrex. It also established the fact that wollastonite was particularly adapted for use in the manufacture of mineral wool, towards which further research was then directed. Briefly, the prin¬cipal advantages were the substantially unvarying chemical composition of the naturally occurring mineral and its characteristic of being relatively flexi¬ble and highly resilient when fused into a glass. A mineral wool produced therefrom is thus not only of highly uniform composition and color but is most desirable because of the high mechanical strength and length of the individual fibers constituting the wool. It is well known that mineral wool may be prepared by melting a suitable charge of raw material and then intro¬ducing a stream of a suitable fluid, such as air or steam, into a stream of the molten raw material. By this means the molten material is blown into threads or fibers, which are collected in a suitable chamber similar to a tun¬nel, from which they are removed from time to time or continuously. Com¬mercial manufacture of mineral wool has been carried on for only about 50 years, but volcanoes, specifically Kilauea in the Hawaiian Islands, have

Jan 1, 1936

Coal - Surface Mining Coal Preparation Coal Utilization Research and Development Health and Safety

Jan 3, 1976

Discussion of the paper of Archer E. Wheeler and Milo W. Krejci, presented at the Butte meeting, August, 1913, and printed in Bulletin No. 80, August, 1913, pp. 1831 to 1880. BRADLEY STOUGHTON, New York, N. Y.:-I would like to ask Mr. Krejci whether they have tried mixing coke dust with the lime in those converters and, if so, if any advantage has been found. It has been a very great advantage in the steel converters where coke dust has been mixed with the lime near the tuyeres. Of course the agent that causes the corrosion at that point is oxide of iron, and coke seems to prevent the formation of the oxide of iron and so protect the tuyeres. It would be interesting to know if it had the same effect in the copper converters. MILO W. KREJCI:-We have never tried it here. At Anaconda the tuyere section was formerly put in by mixing magnesite with tar and ramming that into the tuyere section, but they have abandoned that. Might I ask how you would apply the coke dust? How get it into the tuyere section ? How hold it there ? BRADLEY STOUGHTON :-The coke dust is mixed with silica and clay and then rammed in and burned. I suppose in your converters it would be mixed with the magnetite or magnesite and then agglomerated with the material in place. MILO W. KREJCI:-We never tried it here. PROF. JOSEPH W. RICHARDS, South Bethlehem, Pa.:-I would like to ask how thick this deposit of magnetite is on the lining. How thick a layer would be practical ? MILO W. KREJCI:-About 21 in. would be sufficient. SOLOMON LE FEVRE, Mineville, N. Y.:-I was interested to hear of the use of magnetite in connection with the copper smelter. I am interested in the mining of magnetite, but was not aware that it had been used at all' by the copper metallurgists.

Jan 11, 1913

R. W. Raymond, New York, N. Y.: Mr. Bain's introductory statement that there is a growing disposition on the part of the public " to change emphasis " in government from political equality " to " social justice " is somewhat vague; but I gather from his argument that he approves the growing disposition to which he refers, and that the " emphasis " which it demands involves a radical change in the policy hitherto pursued by the United States—the policy under which the Pacific States and Territories have been developed thus far. I do not understand how "social justice" is to be furthered by the sacrifice of " political equality," or how the change is to be effected by "emphasis." But I fancy that the beneficent and necessary

Jan 1, 1913

By J. Nagy, D. W. Mitchell, E. M. Murphy

In 1957 research was initiated in the U.S. Bureau of Mines experimental coal mine near Pittsburgh, Pa., to study factors affecting foam generation and transport, to evaluate the effectiveness of high-expansion foam for controlling mine fires, and to develop techniques for applying the method under U.S. mining conditions. These investigations showed that high-expansion foam containing at least 0.2 oz of water per cu ft of foam is effective in controlling experimental underground fires burning coal, wood, and oil. Sometimes the fire was completely extinguished, but more often, it was brought under sufficient control to permit either a direct attack on the fire with a stream of water or loading of the hot material into cars. A progress report' prepared in July 1958 summarized the initial achievements of the USBM experiments. Since then other phases of the foam-plug method for attacking fires have been studied in the laboratory and in the mine. Previous studies by British engineers' of the foam-plug method for fighting mine fires indicated that high-expansion foam was effective in controlling experimental timber fires in an underground passageway. Their subsequent workx-1 pertained to the practical aspects of fighting large fires within a mining area with a foam-plug. CONTROLLING EXPERIMENTAL FIRES In the USBM tests foam was formed by spraying a dilute solution of a foaming agent on a metal or cotton net of 1/8 to 1/4-in. mesh. Air passing through the continuously wetted net forms bubbles of 1/2 to 11/2-in. diam and produces a honeycomb of foam that fills the passageway. Under the ventilating-air pressure, this light-weight plug moves forward through the passageways, around sharp corners, and over obstacles. as illustrated in Fig. 1. High-expansion foam was transported to a wood fire, an oil fire, and 13 coal fires. Figs. 3 and 4 show a typical coal fire before and after attack with foam. In 12 of the 15 experiments the fire was brought under control when the water content of foam was 0.2 oz or more per cu ft. A fire was considered controlled when the flames were quenched and observers could cross the area without wearing breathing apparatus or protective clothing. In the other three experiments, conducted when the water content was less than 0.2 oz per cu ft of foam, the flames were retarded but the fire was not controlled. Coal fires have been attacked successfully by foam introduced at points varying from 155 to 1010 ft from the fire. The time of burning in coal beds 10 in. thick ranged from 11/2 to 5 hrs or more. Most of the experimental fire beds were 15 ft in length. However, in one experiment a floor fire 25 ft long and 5 ft wide was constructed $5 upwind from another fire 15 ft in length; in another instance, the fire was 100 ft long and 5 ft wide. Foam was applied to the fires for periods ranging from 7 to 36 min. The time required for foam application depends on the extent of the fire, time of burning, water content of foam, foam velocity, and degree of fire control desired. In addition to the coal fires, foam was transported to a fire covering 45 sq ft, produced by 15 gal of oil burning in metal trays on the floor. The foam extinguished the oil fire in about 1 min. In one other test, the burning of 1100 lb of dry sawmill slabs stacked in open cribs 4 ft high and 16 ft long was brought under control by foam in 2 min. Composition of Gases in Return Air: In several of the experiments samples of the return air from fire zones were collected; composition of the atmosphere before, during, and after foam application was then determined. Because of condensation in the relatively cool sampling tube, the amount of water vapor was not determined. Analyses showed that concentration of carbon dioxide and combustible gases increased as the foam began passing over the fire. This resulted from the decrease in the volume of air when foam generation started and from the formation of gases when water reached the fire.* The quantity of gases generated would not be greater than that from an equivalent amount of water applied directly to the fire. The highest total concentration of combustibles (CO, CH1, and H2 mixture) obtained during the experiment was about 2 pct; this occurred 6 min after foam reached the fire. This atmosphere was nonex-plosive, but calculations show that if the air flow were reduced to about 5 fpm and if the rate of gas liberation from the fire remained constant, the mixture would be explosive. The use of foam on a fire in all probability would affect the normal ventilation of a mine. If the mine is gassy, this factor must be carefully considered before the foam is applied. APPLICATION OF THE FOAM-PLUG TECHNIQUE IN MINES Equipment and procedures for applying the foam-plug methods must be adapted to the prevailing conditions at a particular mine. Some factors to be considered in developing equipment are: size or extent of the mine, dimensions and number of entries, ventilation system, mining methods, haulage facilities, availability of water, amount of methane liberated, and existing fire-control apparatus. • In most experiments the initial air velocity of 200 fpm decreased to 50 to 100 fpm as the foam plug increased In length.

Jan 1, 1961

Book Notices CATALYSIS IN INDUSTRIAL CHEMISTRY. By G. G. Henderson. Lond. and N. Y., Longmans, Green and Co., 1919. 202 pp., cloth, 9 X 6 in., $3.25. This, the latest volume of the series of Monographs on Industrial Chemistry which is being published under the editorship of Sir Edward Thorpe, is intended to present broadly, our present knowledge of the phenomenon of which it treats. The hook reviews the existing literature on catalysis, paying particular attention to the many applications now used in industry. COMMERCIAL OILS-VEGETABLE AND ANIMAL. With special reference to Oriental Oils. By I. F. Laucks. 1st edition. N. Y., John Wiley and Sons, Inc.;, Lond. Chapman and Hall, Ltd., 1919. 138 pp., cloth, 8 X 5 in., $1.25. Intended to give to men in the oil trade, who have not had technical training, the technical data and information required in every-day dealings, in concise form. The various commercial oils are listed, together with their sources, uses, and phys-ical and chemical characteristics. The values have been collected from the standard texts, supplemented, particularly in regard to oriental oils, by the results obtained in the author's laboratory. FIFTY YEARS OF IRON AND STEEL. By Joseph G. Butler, Jr. Cleveland, The Penton Press, 1919. 145 pp., por., 10 X 6 in., leather. (Gift of author.) This volume is one of a special edition, of 100 copies, of an address delivered before the American Iron and Steel Institute in 1917, in which the author gives his personal reminiscences of the development in iron and steel making, as observed during an active participation in the industry in eastern Ohio, which began in 1857. To this are added some brief data on the early history of iron and steel, and a chapter on the activities of the American iron manufactures during the great war. The volume is profusely illustrated with portraits of men of note in the industry.

Jan 12, 1919

By Raymond Mancha

TO determine the optimum inside dimension of an airshaft, it is necessary to strike the proper balance between the cost of power for air friction and turbulence losses within the airshaft, on the one hand, and the cost of sinking the airshaft, on the other. An airshaft larger than necessary is extravagant because the increased amortization rate exceeds the reduction in power cost. Conversely, an airshaft smaller than is needed wastes more power than can be justified by the corresponding reduction in amortization rate. The airshaft power is a function of the air volume Q cu. ft. per min. and the air-pressure drop p in. water gauge across the airshaft. † The airshaft pressure drop is calculated by the formula: R X l X O X Q2 P= 5-2 X A3 wherein R is the air-friction coefficient for the portion of the airshaft 1 ft. deep, 0 ft. in perimeter and A sq. ft. in area, when passing an air volume of Q cu. ft. per minute. The airpower w kw. dissipated within the airshaft is computed by the formula: w= 0.746Q X P 6345 With a combined over-all unit efficiency of the fan and drive of E per cent, the air-shaft power requirements become W = 100w There are 8760 hours per year, so with energy at I cent per kilowatt-hour the annual cost per kilowatt for continuous operation is $87.60; therefore, with energy at C¢ per kw-hr. the annual airshaft power cost MI dollars per year becomes: M1 = 87.60W X C which, expressed by the component parts, becomes: M1 V 5.2X4' / XQ Z 0.746 X 100 X 87.60 X C V 60 X E ) The amortization rate a dollars per year per dollar invested includes both depreciation and interest on the investment. These in turn involve the expected useful life of the airshaft and available molletary interest rate. For the purpose of this discussion the cost of airshaft sinking is assumed to comprise a fixed move-in charge Y dollars, which covers the cost of transporting, erecting and removing the shaft-sinking equipment and is independent of the air-shaft size. Also, it is assumed that the cost of shaft lining, if any, is fixed at $X per cu. yd. of lining material and that excava-

Jan 1, 1947

By Raymond Mancha

TO determine the optimum inside dimension of an airshaft, it is necessary to strike the proper balance between the cost of power for air friction and turbulence losses within the airshaft, on the one hand, and the cost of sinking the airshaft, on the other. An airshaft larger than necessary is extravagant because the increased amortization rate exceeds the reduction in power cost. Conversely, an airshaft smaller than is needed wastes more power than can be justified by the corresponding reduction in amortization rate. The airshaft power is a function of the air volume Q cu. ft. per min. and the air-pressure drop p in. water gauge across the airshaft. † The airshaft pressure drop is calculated by the formula: R X l X O X Q2 P= 5-2 X A3 wherein R is the air-friction coefficient for the portion of the airshaft 1 ft. deep, 0 ft. in perimeter and A sq. ft. in area, when passing an air volume of Q cu. ft. per minute. The airpower w kw. dissipated within the airshaft is computed by the formula: w= 0.746Q X P 6345 With a combined over-all unit efficiency of the fan and drive of E per cent, the air-shaft power requirements become W = 100w There are 8760 hours per year, so with energy at I cent per kilowatt-hour the annual cost per kilowatt for continuous operation is $87.60; therefore, with energy at C¢ per kw-hr. the annual airshaft power cost MI dollars per year becomes: M1 = 87.60W X C which, expressed by the component parts, becomes: M1 V 5.2X4' / XQ Z 0.746 X 100 X 87.60 X C V 60 X E ) The amortization rate a dollars per year per dollar invested includes both depreciation and interest on the investment. These in turn involve the expected useful life of the airshaft and available molletary interest rate. For the purpose of this discussion the cost of airshaft sinking is assumed to comprise a fixed move-in charge Y dollars, which covers the cost of transporting, erecting and removing the shaft-sinking equipment and is independent of the air-shaft size. Also, it is assumed that the cost of shaft lining, if any, is fixed at $X per cu. yd. of lining material and that excava-

Jan 1, 1947