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By Robert Franke

(Butte Meeting, August, 1913.) IN view of the prominence which the conical mill has attained in the fine-crushing field .within the few years since its introduction, the following comparison with its more mature forerunner, the Chilean mill, based on extensive tests, is submitted in the interest of the milling profession. Soon after designing the concentrating plant of the Miami Copper Co. in 1909, the Hardinge conical mill made its appearance in the milling forum. . Its possibility as a suitable crushing device for the plant was well recognized, but in view of the lack of commercial demonstration, at the time, as to capacity, efficiency for desired product, and the still more uncertain factors of cost of maintenance and power consumption, it was deemed that the immediate adoption of this machine throughout the plant would be a hazardous undertaking. For these reasons it was decided to equip the majority of the immediately required units of the plant with Chilean mills, the fine-crushing proficiencies of which were better known, and one section with Hardinge mills, to serve as a test unit for the guidance of future installations and replacements. Thereby, after 1.5 years' operation with both types of mills, a thorough test as to metallurgical efficiency and cost economy has been obtained. The conical mill used in these tests is the 8-ft. Hardinge pebble mill, having a cylinder 22 in. in length. The cylindrical portion of this mill is lined with cast-iron liner plates, and the conical extensions with silex bricks bound together by cement. Each. liner plate carries a projecting lifter, the function of which is to increase the height of drop of the lifted material. Danish No. 5 pebbles, obtained from the coast deposits of Denmark, are used for the grinding charge. The Chilean mill used is a fast-running, 3-roller, 6-ft. Saturn mill, with screens of 0.037-in. opening. The feed to these mills is the oversize of Callow screens having 0.029-in. openings, which. follow rolls crushing to 0.5 in.

Jan 7, 1913

By R. P. Rothwell

The milling of silver ores has arrived at a great degree of perfection in the mining districts of our Western States and Territories, and I hare thought the record of the practical results obtained at the present time, in a few of the principal districts, would prove of value arid interest to many of our members, both at home and abroad. An American silver mill is composed, in descending order from where the ore is dumped, of: 1st. A rock-breaker, usually of the Blake or Alden pattern; 2d. The battery of stamps, the number of which varies from 5 to 120, generally in batteries of 5 or 10 heads, so arranged that a single battery may stand while the others are running; 3d. The tanks where the crushed ore settles (if it is a met stamping-mill), or of a drying-floor if the crushing is done dry; 4th. Amalgamating pans ; 5th. Settling pans ; 6th. Agitators or pans in which mercury escaping in the tailings is caught; 7th. Blanket sluices, over which the tailings run, and where mercury and unamalgamated sulplides are saved. A reference to the engravings (see plate), which show one of the latest forms of an 80stamp silver-mill run by a turbine water-wheel, will show clearly the arrangement of the several parts. THE ONTARIO, UTAH, SILVER MILL. One of the best examples of a silver-mill, and at the same time one of the most successful concerns in the West, is the mill of the Ontario Silver Mining Company, near Salt Lake, in Utah. This mill treats the ore from the Ontario mine, ore which at present is very base, being composed of zinc, lead, and silver sulphides and silver chloride in a quartz gangue. This ore has become baser as the mine attained greater depth, though the vein holds its own, or rather increases, both in thickness and richness. The mill is managed in a skilful and economical manner, though the necessity for

Jan 1, 1880

By Glenn A. Keep

THIS paper is not an announcement of the successful conclusion of the Cerro de Pasco Copper Corporation's pacos-pyrite problems, but merely a description of the commercial-scale, intermittent-roasting operations conducted by the writer, J. H. Lucas and others of the staff. It was prepared by request, for presentation at the February meeting of the Institute. The results obtained in Peru confirm the 1-ton-a-day experiments completed about three years ago at Kearny, N. J. by C. C. Gordon and Mr. Lucas, where it was first demonstrated that a mixture of about 80 per cent. pacos ore with 10 per cent. pyrite ore and 10 per cent. salt could be roasted intermittently in a small 1-ton-a-day Holt-Dern furnace without extraneous fuel, producing porous calcine that could be leached rapidly. Mr. Keep and his staff, by successfully introducing improvements in the construction of the commercial furnace and in their handling of material, have established beyond doubt a low cost for roasting. The other steps in the process of treating pacos and pyrite ores have not as yet been demonstrated as sufficiently conclusive to warrant their acceptance for a permanent plant. Cerro de Pasco Ores The ores of the Cerro de Pasco Copper Corpn. at Cerro de Pasco, Peru, are situated at an altitude of 14,000 ft. The town and mines are served by the Cerro de Pasco Railroad, which operates between Cerro de Pasco and Oroya, the site of the company's copper and lead smelters. Table 1.—Average Analysis of Cerro de Pasco Ores Pacos Ore Pykite Ore Silver, oz............................................... 10.3 12.0 Gold, oz................................................ 0.007 0.01 Lead, per cent........................................... 1.7 1.8 Copper, per cent......................................... 0.19 0.27 Zinc, per. cent........................................... 1.4 1.02 Iron, per cent........................................... 13.4 34.45 Silica, percent.......................................... 61.8 24.5 Alumina, per cent....................................... 2.5 0.62 Lime, per cent........................................... 1.5 0.20 Manganese, per cent..................................... 0.42 0.09 Sulfur, per cent,.......................................... 1.1 34.2

Jan 1, 1929

By Clarence Carleton Semple

At a number of mines in eastern Nicaragua, 3.5- or 5-ft. Huntington mills are used for grinding gold-ore after a pre-liminary breaking in jaw-crushers. The smaller mills are made sectional to facilitate transportation to the more difficultly accessible mines. The capacity of the smaller mills is about 20 or 25 per cent. that of the larger, but in proportion to the quantity of ore ground they require more power and are subject to greater wear. The notes herein given are from records of the performance of the larger mills at one of the mines, but, in general, apply to the smaller mills also. The Nature of the Ore. The ore occurred as a series of flat veins in a' highly-decomposed rock, probably alaskite. The primary mineralizatioll comprised pyrite, pyrrhotite, marcasite, chalcopyrite, and quartz; the oxidation-products of which were chiefly hydrated iron oxide or limonite, chrysocolla, azurite, and malachite. The kaolinization of the feldspar of the country resulted in the admixture of a great quantity of clay with the oxidized ore. The quartz was sandy, imbedded in the clay; and probably the greater part represented the residue from the crushed rock in the fissures after leaching, and kaolinization of the feldspar. There was little vein-quartz in the unaltered primary sulphide ore. Only the oxidized ore was mined for the gold that it contained in amouiits that averaged $4 per ton. No specks of gold mere ever seen in the ore, and the gold obtained by panning was always exceedingly fine. The amalgamation-bullion was about 800 fine in gold and from 60 to 80 in silver. The ore delivered at the mill contained from 5 to 15 per cent. of moisture; during the rainy season the quantity was much greater. The moisture and clay made the ore so sticky that it had to be raked over steeply-inclined grizzlies; it hung in the ore-bins,

Jan 1, 1912

By John S. Marshall

The purpose of this paper is to present to the profession data and experience obtained over a period of 5 years in the operation of 87 shaker-conveyor units, and the production of 2,169,638 tons of run-of-mine anthracite coal. The data are not generally available and should be of interest to those concerned with the operation of anthracite and bituminous mines. To simplify comparison, a ton of 2000 lb. of raw run-of-mine coal has been used as the production unit throughout this paper. Mining Methods and Equipment The mine is in the vicinity of Wilkes-Barre, Pa., in the Buried Valley area of the Susquehanna River. Seven veins were mined, varying in thickness from 18 in. to 8 ft. The cover overlying the veins varied from 200 to 800 ft. Three-fourths of the 1200 acres of property was covered with from 100 to 200 ft. of water-bearing gravels and clays, which prevented first mining under rock cover less than 50 ft. and pillar removal under rock cover less than 100 ft. The property was opened by two shafts sunk on the mountainside to avoid the sand and gravel present in the river valley. The veins dipped an average of 5" from the shafts. In certain areas the veins were practically flat and in some places a reverse pitch as high as 5" was encountered for short distances. A daily output of 3800 tons was hoisted on a two-shifi; hoisting operation. The mining consisted of 75 per cent pillar removal, and of the total production in the later years of operation 80 per cent was loaded by shaker conveyors, chain conveyors and scraper loaders. Shaker conveyors accounted for 70 per cent of the total output; chain conveyors and scraper loaders for 10 per cent. All conveyor work was double-shifted and 25 per cent was triple-shifted. A total of 87 shaker-conveyor units were on the property, of which 75 were kept in operation at all times.

Jan 1, 1940

By John S. Marshall

The purpose of this paper is to present to the profession data and experience obtained over a period of 5 years in the operation of 87 shaker-conveyor units, and the production of 2,169,638 tons of run-of-mine anthracite coal. The data are not generally available and should be of interest to those concerned with the operation of anthracite and bituminous mines. To simplify comparison, a ton of 2000 lb. of raw run-of-mine coal has been used as the production unit throughout this paper. Mining Methods and Equipment The mine is in the vicinity of Wilkes-Barre, Pa., in the Buried Valley area of the Susquehanna River. Seven veins were mined, varying in thickness from 18 in. to 8 ft. The cover overlying the veins varied from 200 to 800 ft. Three-fourths of the 1200 acres of property was covered with from 100 to 200 ft. of water-bearing gravels and clays, which prevented first mining under rock cover less than 50 ft. and pillar removal under rock cover less than 100 ft. The property was opened by two shafts sunk on the mountainside to avoid the sand and gravel present in the river valley. The veins dipped an average of 5" from the shafts. In certain areas the veins were practically flat and in some places a reverse pitch as high as 5" was encountered for short distances. A daily output of 3800 tons was hoisted on a two-shifi; hoisting operation. The mining consisted of 75 per cent pillar removal, and of the total production in the later years of operation 80 per cent was loaded by shaker conveyors, chain conveyors and scraper loaders. Shaker conveyors accounted for 70 per cent of the total output; chain conveyors and scraper loaders for 10 per cent. All conveyor work was double-shifted and 25 per cent was triple-shifted. A total of 87 shaker-conveyor units were on the property, of which 75 were kept in operation at all times.

Jan 1, 1940

By H. l. Burghoff, D. E. Lawson

The presence of tellurium in copper and, by inference, in copper alloys, has been considered seriously detrimental and has been avoided. In particular, very small amounts of tellurium have been found to produce red-shortness, although the effects on cold-working and ordinary physical and mechanical properties have not been known so definitely. References in the literature are few, but as long ago as 1881, Eglestonl reported that 0.083 per cent tellurium in fire-refined copper rendered the copper red-short and therefore worthless for hot-rolling. In hot-rolling copper cakes containing this amount of tellurium, he found that fine cracks appeared in the cakes on the first pass through the rolls, the cracks opening further in successive passes. When cold, the metal was said to be tough and malleable. A few years ago, during the course of an investigation in our laboratory on copper-base alloys, it was found that tellurium-bearing alloys could be machined much more freely than similar alloys not containing tellurium. The study was pursued further in this direction, and considerable data were obtained, which indicated that certain tellurium-bearing alloys might have properties of commercial interest. The work eventually led to patent protection,2 a wide range of tellurium-bearing copper-base alloys being covered. This article presents results obtained on the effect of tellurium on the mechanical properties of three copper-base alloys: commercial bronze (90 Cu, 10 Zn), 1.5 per cent tin bronze (98.5 Cu, 1.5 Sn), and silicon bronze (96 Cu, 3Si, 1 Zn). Several alloys were cast in each series, with tellurium content ranging from 0 up to about 1 per cent. The alloys were melted under charcoal in a clay-graphite crucible, using a high-frequency induction furnace, and were cast without undue difficulty in rod form, 15/8 in. in diameter. Charges for the zinc and tin series were made up from the following materials: cathode copper, electrolytic zinc, tin, and tellurium in slab

Jan 1, 1938

By H. l. Burghoff, D. E. Lawson

The presence of tellurium in copper and, by inference, in copper alloys, has been considered seriously detrimental and has been avoided. In particular, very small amounts of tellurium have been found to produce red-shortness, although the effects on cold-working and ordinary physical and mechanical properties have not been known so definitely. References in the literature are few, but as long ago as 1881, Eglestonl reported that 0.083 per cent tellurium in fire-refined copper rendered the copper red-short and therefore worthless for hot-rolling. In hot-rolling copper cakes containing this amount of tellurium, he found that fine cracks appeared in the cakes on the first pass through the rolls, the cracks opening further in successive passes. When cold, the metal was said to be tough and malleable. A few years ago, during the course of an investigation in our laboratory on copper-base alloys, it was found that tellurium-bearing alloys could be machined much more freely than similar alloys not containing tellurium. The study was pursued further in this direction, and considerable data were obtained, which indicated that certain tellurium-bearing alloys might have properties of commercial interest. The work eventually led to patent protection,2 a wide range of tellurium-bearing copper-base alloys being covered. This article presents results obtained on the effect of tellurium on the mechanical properties of three copper-base alloys: commercial bronze (90 Cu, 10 Zn), 1.5 per cent tin bronze (98.5 Cu, 1.5 Sn), and silicon bronze (96 Cu, 3Si, 1 Zn). Several alloys were cast in each series, with tellurium content ranging from 0 up to about 1 per cent. The alloys were melted under charcoal in a clay-graphite crucible, using a high-frequency induction furnace, and were cast without undue difficulty in rod form, 15/8 in. in diameter. Charges for the zinc and tin series were made up from the following materials: cathode copper, electrolytic zinc, tin, and tellurium in slab

Jan 1, 1938

By Frank Firmstone

The piece of spiegeleisen, the analysis of which is given below, was collected by my father, together with various other specimens, while he was manager of the Glendon Iron Works. It bears a label stating that it was taken from a building in the town of Newton, New Jersey, erected in 1765 and torn down in 1863. The building belonged, in 1863, to T. N. McCarter, Esq. The label is in the hand-writing of Mr. McCarter, now deceased, well known as a member of the bar of northern New Jersey. The specimen was, no doubt, given by him to illy father, with whom he had business in 1863 and 1864. It is now in the metallurgical collection of Lehigh University. I know nothing further as to the history of the specimen, but there is no reasonable doubt that it was made from ore from the Andover mine, at the old charcoal-furnace which once stood at the village of Andover, about 5 miles south of Newton, on the Sussex railroad. It was probably broken from one end of a bearer over an open fire-place, is about 6 in. wide by 2.5 in. thick, has been cast in open sand and shows the charcteristic fracture of spiegeleisen rather low in manganese, as shown in Table I. One edge has been white-washed. Table I.—Analyses of Spiegeleisen Made from Andover Ore. 1. 11. Per Cent. Per Cent Graphite,.0.101 0.005 Combined carbon,... 4.529 4.212 Silicon,..0.037 0.294 Manganese,..... 5.760 3.750 Sulphur,.0.003 0.031 Phosphorus,.0.060 0.072 Iron,........ 89.440 91.570 99.920 99.934 I. Spiegeleisen made at Andorer (charcoal) furnace not later than 1765. IT. Spiegeleisen made from Andover ore at Phillipsburg, N. J., anthracire-furnwe, 185-. Both analyses were made by Dr. P. W. Shimer.

Jan 1, 1907

By C. W. Thompson

Paul D. Suloff (Goodyear Tire and Rubber Co., Inc., Akron, Ohio)—I would like first to comment on problems of the conveyor belt discussed in Mr. Thompson's excellent paper, since that is what we hope we know most about. Twists in relatively wide conveyor belt unavoidably produce a lateral maldistribution of tension, raising tension at belt edges and reducing it at the center. They also produce a lateral collapsing force on the belt at the center of the twist owing to the inherent tendency of all the longitudinal elements of the belt to try to pass through a point at the twist center. Calculation of the twist geometry by the methods shown in Mr. Thompson's paper keeps these extraordinary forces within limits which the belt designer can tolerate. No reduction in belt life due to twisting need be contemplated when this geometry is maintained. There is a minor exception that belts of extreme lateral flexibility will tend to curl laterally at the center of the twist. However, any ordinary fabric construction will perform satisfactorily in this respect. These twists are always made in regions of low tension in the conveyor so that even in the edges of the twist, belt tension does not exceed the average tension found in highly stressed regions of the conveyor. Offsetting these out-of-ordinary belt stresses is the advantage that Mr. Thompson has brought out of getting the return run up out of the dirt where it can be seen. This not only makes it easier to train, but also, in the event that it is not properly trained, frees it of the normal return run edge wear hazard. It is well known that return run edge wear is a prominent cause of belt mortality underground. An interesting aspect of this two-way conveyor is that the belt may be made what is known as a Mobius Strip. A Mobius Strip is obtained by splicing a belt after turning one end of it 180" about its longitudinal axis. In other words, one end is turned upside down before splicing. A belt spliced in this fashion turns itself upside down every time it comes around, but the twist which has been put in the splicing, of course, stays at one location on the conveyor, in this case one of the twist sections at the end. Turning the belt over every revolution might have advantages in some cases. Belts could be made with equal covers and the two sides worn equally and simultaneously. In this case there would be no problem of getting belts on upside down by mistake. However, the two-way conveyor does not have to be a Mobius Strip. It can be twisted in such fashion that the same side is up on both runs. It is simply a question of which way the final 90" twist is made before joining the ends. Another interesting aspect of the two-way conveyor is the problem of operating two-way conveyors in series. Here the sequencing of starting brings up some new problems. It will be recognized, although not always at first glance, that if the starting sequence is planned for one run of the conveyor the reverse will result on the other run. With the two runs carrying bulk material in both directions a reverse sequence on one run would be intolerable. In this situation the only solution appears to be a simultaneous starting of all conveyors in the series. However, with the coal in one direction and intermittent supplies in the other it would be entirely practical to sequence the conveyors for the coal run and accept a reverse sequence on the supply run. The two-way conveyor also lends itself to new driving possibilities. First, it is quite possible to drive at the head end of each run, which of course, means a drive at each end of the two-way conveyor. Driving in this way a given belt can be extended to substantially greater lengths than a conventional conveyor with drive at one end only. In addition to this, under certain conditions the conveyor can be extended to extreme length by driving at one end and at some intermediate point on the most heavily loaded run. As a particular case, a belt carrying coal downgrade and supplies back upgrade could be extended to extreme lengths by driving at the head of the coal run and at an intermediate point of the supply run. Mr. Thompson has been a pioneer in belt conveyor transportation underground and his accomplishment here with the first two-way conveyor of any consequence is another notable addition to the art.

Jan 1, 1954

By Robert M. Dreyer

Diamond is the hardest known material. The diamond industry is separated into two major segments: (1) industrial and (2) gem. The major industrial use of diamonds is as a high-grade abrasive in a wide variety of applications for which no effective low-cost substitute is available. Table 15.4C.1 gives data on diamond production and on United States diamond imports. In both gem and industrial diamond production, but especially in the case of gems, the quantity of stones produced is not directly correlated to value, for the determination of value is markedly affected by quality. South Africa long has been the world leader in gem diamond production with an estimated production in 1968 (including South-West Africa) of about 4 million carats, equal to 43% of world production. During the past decade there has been a marked rise in gem diamond production in the USSR, so that nation is now the second-ranking world producer, with Angola (Portuguese West Africa) third. In natural industrial diamonds, the major production comes from Congo-Kinshasa (former Belgian Congo) which had a production of 15 million carats in 1968, equal to 47% of world production. In the past six years, there has been a remarkable increase of industrial diamond production in the USSR, which now ranks second with a production of 6 million carats, equal to about one-fifth of world production. South Africa now ranks third followed by Congo-Brazzaville and Ghana. No natural diamonds are produced in the United States. Since the discovery of an economical method for synthesizing small diamonds, there has been a rapid increase in synthetic production from 4 million carats in 1962 (1 carat equals 200 mg) to 11.5 million carats in 1968. Table 15.4C.1 gives the gradually increasing imports of industrial diamonds and indicates a relatively stable price. In 1968, the imports of industrial diamonds were 13.5 million carats with a value of $59.2 million. Table 15.4C.2 shows the rapidly increasing imports of cut and uncut gem diamonds into the United States. In 1968 the United States imported 2.5 million carats of uncut stones with a value of $166 million and 1.5 million carats of cut stones with a value of $175 million. Diamond marketing is dominated by DeBeers Consolidated Mines, Ltd., which, through subsidiaries, controls many of the major African diamond deposits. More important economically, however, through its Central Selling Organization, DeBeers markets 86% of the non-Communist world's natural diamonds, including most of those derived from deposits outside of corporate control. In 1967 the mines owned by DeBeers produced nearly 8 million carats, and the Central Selling Organization sold over $600 million in gem and industrial diamonds. The dominance of diamond sales by

Jan 1, 1976

DeLamar, the largest silver mine in the United States outside the Coeur d’Alene District in Idaho, came into production in April 1977. It is a joint venture of Earth Resources Company and Canadian Superior Oil Co., with Earth Resources being the operator. The property is a consolidation of several old operations dating back to the 1800s; the better known of these are the DeLamar and Sommer-camp mines. DeLamar is situated in southwestern Idaho near the Oregon border in the Owyhee Mountains, 88 km (55 miles) southwest of Boise, Idaho. Heavy snowfalls and very low temperatures are common in the wintertime. Planned production was 77,500 kg (2.5, million oz) of silver and 5 13 kg (16,500 oz) of gold per year. Expected life at a milling rate of 1,540 mtpd (1,700 stpd) was 20 years, with an ore grade of 144 g Ag and 1.58 g Au per mt (4.2 oz Ag and 0.046 oz Au per st). Present production is at the rate of 1,900 mtpd (2,100 .stpd). The total cost to bring the mine into production was reported to be approximately $22 million, and the mill was reported to have cost about $10,000 per daily ton of design capacity. At DeLamar, naurnanite (Ag2Se) is the principal silver .mineral occurring with acanthite (Ag2S) disseminated through quartz veins and a silicified rhyolitic ash-flow tuff of Miocene age. The mineralization is relatively iron and manganese free and is readily amenable to cyanidation giving recoveries of 85% silver and 92% gold, which occurs free. When the plant was started, considerable difficulties were encountered in handling the wet sticky ore because of the high clay content. Following primary crushing, the ore is stockpiled. Due to the sticky nature of the ore, the vibrating feeders initially installed under the stockpile were replaced by a reclaim hopper and an apron feeder which was fed 24 hours a day by a front-end loader. The minus 200 mm (8 in.) ore is fed to a 5.5 m by 2.75 m (18 ft by 9 ft) autogenous mill followed by a 2.75 m by 4.57 m (9 ft by 15 ft) ball mill and cyclones. The mills are equipped with 750 kw and 520 kw (1,000 hp and 700 hp) motors. Both mills discharge to the same cyclone feed sump and a portion of the cyclone underflow is returned to the autogenous mill. Cyclone overflow, which is 20% plus 74 microns (200 mesh), is leached in four agitators in series for 72 hours (45 hours at the 2,100

Jan 1, 1981

By NATHANIEI HERZ

(Cleveland Meeting, October, 1912.) THE formula very frequently used for computing the economical size of pipe to transmit compressed air is that of D'Arcy, as follows: d5(p1-p2)/w1l D-c w11 Where, D = the volume of compressed air delivered in cubic feet per minute,-at the final pressure. c = an experimental coefficient determined for the various sizes of pipe. d = the diameter of the pipe in inches. 1 = the length of the pipe in feet. P1 = the initial gauge-pressure in pounds per square inch. p2= the final gauge-pressure in pounds per square inch. W1 = the density of the air, or the weight in pounds per cubic foot at the initial pressure, p,. The most common case is that in which the given quantities are: the quantity of air required, the length of the pipe, and the initial pressure. The method of solution is to assume a pressure-loss and to compute the remaining factor, c/d5, thus giving the size of pipe corresponding to the assumed loss of pressure. It is always desirable to try two or more pressure-drops, in order to find the combination that is most satisfactory, since often a small change in the size of pipe will reduce or increase the loss of pressure several pounds. An alternative method is to assume a size of pipe and calculate the corresponding pressure-drop. Each method involves a series of tedious calculations to arrive at the most economical solution, and also requires the use of tables giving the constant, c, the actual diameters corresponding to the nominal pipe-sizes, the density of the air, w1, and often, for convenience, a table giving the

Dec 1, 1912

By L. W. Saperstein, W. W. Kaufman

Stimulated by the hearing protection clauses in the 1969 Coal Mine Health and Safety Law, attempts were made to determine what safety hazards, in terms of warning signal discrimination, are attendant upon the wearing of hearing protectors. Warning signals of dangerous roof conditions were collected and analyzed from 11 different coal mines in four different coal seams. Primarily, signals of audible level emanating from roof, floor, or coal were recorded. The interpretations of these signals were seen to be very dependent upon coincident visual and tactile signals. Acoustic character of such signals is shown to be dependent upon the spectral energy distribution of the pulses. Roof talk signals generally have broad-band spectra concentrated below 5 kHz. Reverberation times of acoustic pulses are generally less than 0.5 sec in duration. Indications have developed which link the coal mine entry dimensions with the nature of noises recorded there. Related experiments suggest that roof talk can be discriminated from background noise equally well when wearing protection as when not. This conclusion is only appropriate when noise levels warrant the wearing of protection. When quiet, the wearing of protection somewhat reduces discriminatory ability.

Jan 1, 1976

By Earl Oliver

CONTENTS PAGE A.I.M.E. Stabilization Committee Activity 3 Definition of Capture Rule: Robert E. Hardwicke 4 Westmoreland Natural Gas Co. vs. DeWitt 5 Kelly vs. Ohio Oil Co 5 Bernard vs. Monongahela Gas Co 6 Definition of Ownership in Place 7 Prospective Stabilization Remedies 7 Dean Roscoe Pound's Four Tests of Practicability 7 I: Mechanically Feasible 8 Capture Rule 8 Ownership in Place 10 Mechanical Separation Method 10 Engineering Appraisal Method. 11 II. Economically Advisable. ., . 13 Capture Rule 14 Ownership in Place 14 III. Politically Wise 14 Capture Rule 14 Ownership in Place 17 IV. Legally Sound 18 Capture Rule 18 Ownership in Place 19 Dean Henry M. Bates' Analysis 19 Resume 22 THE A.I.M.E. Stabilization Committee was named in 1932 to ascer-tain whether some practical substitute could be devised for the Rule of Capture, which was generally credited with being the major cause of instability in the Petroleum Industry. After considerable intercommittee correspondence a letter was received from. E. DeGolyer under date of Jan. 25, 1935, in which he discussed the possibility of modifying rather than discarding the Rule of Capture. He summarized his proposal as follows:

Jan 1, 1937

By J. A. Cordier

USINOR has conducted four blast furnace tests with formcoke made from marginal coals in the pilot plant of HBNPC ("H. B. N." process) in four different blast furnaces of various sizes and with medium and high grade burdens. The tests in which up to 30% of the blast furnace fuel burden was replaced by HBN formcoke yielded good results on furnaces ranging in hearth diameters from 5.6 to 9.5 m. The tests demonstrated that: - the slight increase of pressure drop in the stack did not prevent smooth operation at record production rates with the same blowers - the total coke rate was not significantly affected, but there was a slight increase of the calorific value of the top gas - the hot properties of the formcoke are satisfactory as seen from the samples withdrawn at the tuyeres and tapholes. At the end of the two months test on BE nº 2 of Dunkerque, the hearth was as permeable and clean as with good standard coke. Our conclusion is that HBNPC formcoke made from lean coals is an adequate substitute for conventional coke in the range of 20 to 40% replacement, which is of interest to USINOR.

Jan 1, 1977

By R. S. Burns, J. G. Erskine

The performance of the nine 5.5 m diameter x 6.4 m mills (18 x 21 ft) initially installed at Bougainville Copper Ltd. did not meet the laboratory and pilot plant derived design expectations as measured by Bond Work Index. During the post-commissioning period, inefficient grinding of coarse material resulted in a build-up of large circulating loads and led to reduced throughputs because of pump and cyclone overloading. Installation of larger pumps and cyclones and development of an improved circuit control philosophy alleviated materials handling problems and permitted higher tonnages to be attained. Since 1977, plant throughput has been increased by installation of three additional mills. These are 0.9 m (2.9 ft) longer than the other nine and have a higher capacity. Developments in classification and mill speed-up, together with a finer ball mill feed size and substantial improvements to mill availability (99% in 1982) have further increased plant capacity. There is a continuing emphasis on improving grinding circuit efficiency in terms of throughput/grind. Current development work includes trials of graded ball charges, different liner profiles, and assessment of scalping/regrind proposals.

Jan 1, 1985

By C. A. Crawford

NICKEL and the nickel-copper alloy, the latter generally referred to as morel metal, are available in a variety of wrought and fabricated forums, of which the following are regular commercial products: Form Monel Metal Pure Nickel Roc Hot-rolled; round, hexagon, Same as morel metal square, and flat shapes Cold drawn; round, hexagon, square, and flat shapes Hot-rolled wire rod Cold-drawn wire rod Plates Hot-rolled Same as morel metal Sheets Hot-rolled annealed Same as morel metal Hot-rolled annealed, pickled Full-finished Full-finished, No. 3 grind one side Full-finished, No. 4 grind one side Cold-rolled (annealed) Cold-rolled, No. 5 polished finish Strip Hot-rolled (flat rods) Same as morel metal Cold-rolled (annealed) Cold-rolled (unannealed) Tubes and pipe Gas-welded tube Seamless tube Seamless tube in I. P. S. Gas-welded tube Forgings Bolts, rivets Same as morel metal Nuts and washers Forging to order Castings Sand castings Same as morel metal Centrifugal castings Electrolytic Cathodes of high purity Anodes Cast anodes Rolled anodes Shot and block for 25 to 50-1b. pigs 5, 25 and 50-1b. pigs melting. Shot Shot

Jan 1, 1927

By R. Gilman Brown

Mines are interesting by reason of what they have done for man, or of what has been done for them by nature. Not all are interesting on both scores. Many profitable mines are commonplace to the geologist; and many presenting unique gcological conditions have been sad failures commercially. The Standard group at Bodie has produced in 25 years, $14,500,000, paying $5,000,000 in dividends.' On the other hand, this output has beeu mined (almost wholly above the 1,000-ft. level) from a system of more than 100 veins, ranging in thickness from 0.5 in. of ('specimen rock" to 30 ft, of clay and quartz; distributed through a zone about 2,000 ft, in width, and representing from three to five distinct periods of formation. This group, therefore, by reason of its industrial record as well as its geological features, may fairly be regarded as doubly interesting. The Bodie district occupies an island of recent (Tertiary) hornblende-andesite, surrounded by a complex of igneous rocks and breccias. It lies at the summit of the eastern foot-hills of the Sierra Nevada mountains, and within the Great Basin, a

Jan 1, 1908

By Shiou-Chuan Sun

THE effects of oxidation on the flotation behavior of sulphide minerals have been extensively studied,1-3 but no similar study has been made of coals. Coals of bituminous and lower rank undergo atmospheric oxidation during mining and storage, and their degree of oxidizing increases with temperature4,5 and time of exposure.6 The problem of recovering fine oxidized coal from culm piles, rivers, and streams by froth flotation has become increasingly important to the coal industry as well as to those concerned with air and stream pollution. This paper describes the flotation behavior of oxidized and unoxidized coals and carbons in the presence of oily collectors and presents the reagents found effective in the flotation of oxidized coals and carbons. Except where otherwise stated, flotation tests were performed in a 50-g Denver flotation cell. Solid feed for each test was 30 g, and the pulp dilution was kept at 8.5. The -65 mesh coal sample was wetted by agitation with water in the flotation cell for 6 min, then conditioned with a predetermined amount of collector for 2 min, and finally floated with pine oil as frother for 3 min.

Jan 4, 1954