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By John S. Johnson

The maintenance cost, or renewal cost, in any preparation plant is a large item of expense and reflects in the total cost of producing and preparing the ton of coal. It is given constant study to the end that ways and means of controlling and reducing this proportion of the total cost may be developed. Study of this item of cost has proceeded along usual lines, which are: proper functioning of the maintenance organization; systematic and thorough inspection to prevent breakdonns; the making of repairs before renewals are needed; maintenance work done thoroughly the first time, instead of patching, study of parts needing frequent renewal or repairs or conditions resulting in delays, thorough knowledge of cost by the mall responsible for maintenance; and the use of special equipment and materials. This paper is to deal with special equipment and materials to reduce the maintenance cost hut before going into that I want to express an opinion on the general subject. Maintenance Organization Usually maintenance is taken care of by a group of men who perform their work on the off shift. The work is arranged by the operating official who should be thoroughly familiar with maintenance, so that materials needed may be made up for the particular job and waste of time in looking up materials be prevented. To perfect the maintenance organization, all operating attendants should be thoroughly trained to be familiar with all working parts of their equipment, to make a thorough inspection and to write out the slip for the work required. This requires job analysis, but it is worth the effort, because the attendant usually is better informed than anyone else about his particular equipment. The operating official can judge the scope of the work to direct the repair and to have material ready. Lubrication is an important item in maintenance, and a subject by itself. Usually the breaker oiler is designated to perform this task. This man should be picked for his ability and put in line for promotion to machinist. He should be trained as an inspector, as he is constantly visiting most parts of the breaker and has an opportunity to observe needed maintenance. The first thing is to lay out maintenance work and arrange to have it done; the next is to check the work and see that it has been properly done. This check should be made by the operating official, and through the attendant. The cost statement showing the various items of maintenance cost is not often understood by the man in direct charge. Cost sheets should be simplified and worked up by breaker sections; e.g., the cost of

Jan 1, 1947

By AIME AIME

THOUGH the dust-control systems in the crushing plants and other buildings at Morenci do not differ materially from similar installations in other large copper reduction works, it is probable that in few other plants has as much attention been paid to the dust problem. Possibly the most striking feature of the Morenci installation is the manner in which the necessary equipment has been built into the various structures, and the absence of unsightly pipes, ducts, and fan houses. That this result has been attained is due, in no small part, to the basic conception that the control of the dust in a milling plant should be an essential part of the process itself and not something to be tacked on later after the plant is in operation. The control systems were designed and developed along with the general plant construction program and when the first car of ore was dumped into the crusher the dust-control equipment was ready.

Jan 1, 1942

By John V. Beall

With some difficulty we framed the picture of the village on the mountain getting in the railroad bridge and tunnel at the base. It was a charming scene of red-tile roofs on a green mountain in the golden autumn sunshine with the black hole of the railroad tunnel looking like something out of the movie, "The Train." We clicked the shutter, advanced the film and started back to our rented Audi. At which point there was a hail from the other side of the railroad bridge. A man in uniform was brandishing an object in the air which looked suspiciously like a rifle. We peered around. We were the only other persons in view. He must be yelling at us. He was now closing the distance at a brisk, determined gait but was still 100 yards away. The car was handy but the road was straight, paralleling the railroad tracks for some distance. We dismissed the idea of running for it. Who knows, he might shoot. When he came up we didn't have much of a conversation because it had taken us a week to learn three words of Yugoslavian-zhiveli, long life, huhvalla, thank you, and schlivavitza, the national drink (plum brandy). None of these seemed appropriate. We got the Idea that photographs of railroad tunnels and bridges were prohibited. The guard, for such he seemed to be, took the Audi plate number, administered a last scowling reprimand and stalked off. This incident was the only one in a 15-day visit to Yugoslavia which we would have associated with "Iron Curtain restrictions."

Jan 1, 1970

By AIME AIME

COMMITTEE No. 2 of the Iron and Steel Institute b f Great. Britain has presented its first report, of 27 printed pages, on blast-furnace plant and practice. This report outlines the various features of modern blast-furnace practice with brief comments thereon, and with references to some of the more recent methods devised for applying the principles and attaining the purposes aimed at' under these headings.. In general, the trend in the United States is in the direction of larger and larger furnaces with ever increasing capacity from which, with improved details in methods and equipment, and attention to details of management, considerable progress in. economy of production has been attained, in fuel consumption and in other items of production cost.

Jan 1, 1929

DAVE BARTLEY and Edwin F. Gray, miners, late of Shasta County, California, young, healthy, and "dead broke," arrived at Ely, Nevada, one evening in the summer of 1900. Before rustling jobs and going to work, they decided to look around the camp, and in pursuance of this plan they visited a number of claims owned by D. C. McDonald, Justice of the Peace and pioneer of the district. Among the claims he showed them were the Ruth (named after his daughter) and the Kearsarge. Both were regarded as possible gold or silver properties. For no very definite reason the two miners changed their plan and took a lease and option to buy the two claims for $3500. After the transaction was closed and the partners were alone, Gray asked Bartley what he thought of the ground. "If she's good for anything, it's copper" was the reply; and, so far as the records disclose, that was the first time that anyone ever thought of Ely or the Robinson mining district, as it was called, as a possible copper area. There were some slight surface indications of copper, but it is probable that the fact that Bartley had just come from a copper district had something to do with his predilection for a metal that was much despised around Ely. In any event, his hunch was good, and he spoke with the wisdom of an oracle. For thirty years prospectors had sought gold and silver around the camp. Several stamp mills and cyanide plants had been put up and a hundred shafts had been sunk, but fate had been indifferent in its rewards to the optimistic owners. Copper ore would have to be hauled by wagon, 150 miles down Steptoe Valley to the Southern Pacific Railroad, and there was scant hope of finding ore that would bear this cost. W. B. Graham, who ran the general store in Ely, grubstaked Bartley and Gray, furnishing them with food and mining

Jan 1, 1933

By E. A. Jones

THE Southeast Missouri division of the St. Joseph Lead Co. normally hauls and hoists over 5 million tons of lead ore each year. This ore is mined in the stopes and headings of 20 mines, hauled to a main line system, see above, and transported up to 6.4 miles to ore-hoisting shafts. Average haulage distance is 2.16 miles. A few years ago a distinguished mining engineer visiting this area for the first time remarked, "You do not have a mining problem; yours is a haulage problem." Its relative importance may be gaged by the fact that mine haulage represents about 25 pct of the total cost of mining ore, the most costly single underground operation. The Southeast Missouri lead belt is a mining district located about 65 miles south of St. Louis. This district is roughly 6 miles wide and 10 miles long. The ore horizon lies about 400 ft below the surface where most of the ore is mined from flat-lying beds of dolomitic limestone. This paper deals with the central part of the district, namely, that portion which contains a group of interconnected mines all in a central haulage system, serving four hoisting shafts. The two major hoisting operations are carried out at Federal No. 17 shaft and Leadwood No. 12 shaft. No. 17 accounted for about 70 pct of the ore hoisted in the district during 1950, averaging about 14,000 tons per operating day. The underground dumping and skip-pocket system at this shaft is similar to those of the other shafts except that it has two rotary dumps. The other ore-hoisting shafts have only one. It might be mentioned that waste rock is hoisted at still other shafts, hauled off, and dumped on the surface by trucks. The hoisting shafts are on a connected system of about 200 miles of underground 24-in. gage main line railroad. There is also the Bonne Terre system, with its own ore-hoisting shaft, which consists of 13.5 miles of 24-in. gage main line, located in the extreme north portion of the lead belt area. Smaller outlying mines such as Doe Run No. 12 and Hayden Creek operate on trackless or shuttle car haulage. In addition to the system of main line haulage, each contributing mine group has its own extensive system of stope or heading railroad. Trackless mining is making considerable headway in the Lead-wood and Desloge mines. Here the shovel loads into a shuttle car which in turn dumps into mine cars spotted on the main line track system. When the merits of underground haulage and surface transportation are compared, it must be remembered that in early days there were many individual

Jan 1, 1954

By G. R. Purdy, J. A. Malcolm

The isothermal dendritic precipitation of y in p CuZnSn is compared with the isothermal formation of Widmanstatten a plates in p CuZn. Although both precipitates are evidently bounded by semicoherent interfaces, the dendritic shape of y evolves through smooth "sinuoidal" shape perturbations, while a-p interfaces tend to be strictly planar, or rnicroscopically faceted. An estimate of the average inter-facial free energy of cube-oriented y in p through the Mullins-Sekerka analysis indicates that it is low (of the order of 50 ergs per sq cm). Interface mication is apparently controlled by volunle dij'usion, and the observed anisotropy in growth probably results from anisotropy of inter facial free energy. Studies of the grain boundary precipitation of y in p indicate that the precipitate can grow in both $ grains, but becomes unstable with respect to the grain with which it shares a mobile semicoherent interface. Facets observed at the tips of Kurdjumov-Sacks-oriented Wid-manstatten a plates correspond to variants of an irrational hahit, near (231)0. Comparison with Hu and Smith's determination of relative interfacial free energies of a-(3 boundaries suggests that this system possesses an extremely deep o cup, which results in an ej'ective low-o zone. The possible effects of such a cup on the stability of a-p interfaces are discussed, a needles, grown during rapid continuous cooling, possess microscopically smooth interfaces, and are characterized by a high degree of' damage in the region of the boundaries. These boundaries, induced to move on tempering at 200"C, become faceted as equilibrium is approached, which seems to indicate that boundary structure and mobility in this system are dependent on the local degvee of departure from equilibrium MULLINS and sekerkal have elucidated the phase boundary stabilizing effect of surface tension. a factor not taken into account in previous treatments of morphological development. Their analysis applies to the isothermal growth of a pure spherical precipitate (planar in the limit of infinite radius) in a slightly supersaturated solution, under conditions of inter-facial local equilibrium and isotropy of interfacial free energy. More recently, several other factors which might tend to stabilize planar interfaces have received attention. These are: i) interface diffusion;4"8 and ii) slow interfacial reaction. cahn7 included a slight anisotropy in interfacial free energy, and showed that it has no stabilizing effect, but forces shape perturbations to reflect the anisotropy. shewmon4 noted that dendritic morphologies are rarely observed in solid-solid systems, and, in particular, postulated that slow interfacial reaction occurs at all boundaries between proeutectoid ferrite and austenite, except those which propagate parallel to {11 l)y planes (for Kurdjumov-Sachs [K-S] related phases). Previously, Aaronson~ had proposed interfacial reaction at the broad faces of Widmanstatten plates, and expected that the lateral motion of high-mobility incoherent ledge edges would produce planar precipitates bounded by law-mobility semi-coherent interfaces. The problem of interfacial reaction during the dif-fusional formation of precipitates has often been ignored, although it is apparent that a small component of chemical reaction must exist at any moving interface." Further, all solutions of the diffusion equations for boundary conditions corresponding to isothermal growth predict infinite growth rate at time zero, so that elements of interfacial reaction would be most easily detected at early times. However, reported deviations from diffusion control are comparatively rare," and the local-equilibrium diffusion-control model appears adequate for the description of many isothermal growth systems.'2"9 Cahn14 has treated the problem of migration of an interface which is able to reach positional energy minima, and concluded that such an interface, if subjected to a high local ther-modynamic "driving force", may migrate normal to itself without the benefit of ledges. As local equi-

Jan 1, 1968

By R. S. Kinasevich, D. D. Crabtree, D. W. Fuerstenau, T. P. Meloy, A. L. Mular

The effect of such variables as feed size, particle shape, cleavage, and hardness on the relative proportion of impact, abrasion, and chipping events in a tumbling mill is demonstrated. The concept of an average event, which is a combination of the three comminution mechanisms, is introduced. It is postulated that the average event in a particular mill determines not only the size distribution but also the energy-size reduction relationship for the comminution process. Fragmentation of particles in tumbling mills appears to result from a combination of impact, abrasion, and chipping grinding.' The relative rates of size reduction attributable to any of the basic grinding mechanisms must depend upon a number of variables, such as feed size, particle shape, hardness, and the number and kind of particles in the mill. The purpose of this paper is to show experimentally how these variables in the properties of materials affect the size distributions obtained by grinding the material's in a laboratory tumbling mill. The resulting size distributions will be intetpreted in terms of a combination of three comminution mechanisms. METHODS AND MATERIALS In this study, minerals of widely differing properties were ground either separately or as mixtures in laboratory mills using rods or balls for grinding media. The materials used include quartz, Tuolomne, limestone (99.8% CaCO3), barite, corundum, galena, pyrite, glass beads, and crushed glass. The mill used was 8 in. in diam. and 9-1/4 in. long. For the ball milling tests, 17.6 kg of steel balls (36: 7/8 in. balls, 178: 1 in. balls, and 44: 1-1/8 to 1-1/4 in. balls) were used. For the rod milling experiments, 17.6 kg of steel rods 9-1/16 in. long (25: 5/8 in., 10: 3/4 in., and 5: 7/8 in. in diam.) were used. In each test, 1 kg of solids were ground wet at 60% solids at 77 rpm. In all cases except where otherwise specified, the feed to the mill was 4 x 8 mesh material. Size distributions of the ground products were determined by a wet-dry sieving procedure with Tyler series sieves. Size distributions of mixed products containing limestone were determined by dissolving the limestone with acid and sizing the residue. EXPERIMENTAL RESULTS AND DISCUSSION If the size distribution of the product from a tumbling mill depends upon the relative proportion of impact and attrition (abrasion and chipping) events, altering the grinding conditions may change these proportions and consequently alter the size distribution of the product. Data to illustrate this point can be taken from Gaudin's classic 1926 paper.2 In one series of experiments, Gaudin ground quartz of different initial sizes for four hours in a laboratory ball mill with 7/8-in. balls. In Fig. 1, his data are reproduced as the cumulative fraction finer than the stated size for the following initial feed sizes: 13.3 x 9.4 mm, 6.7 x 4.7 mm, 2.6 x 1.6 mm, and 0.83 x 0.59 mm. As the feed size was decreased, a dramatic change in the shape of the size distribution curves took place, even though the amount of minus 100-mesh

Jan 1, 1964

By John J. Cioston

THE steadily rising flood of gold production from all parts of the world has created an avalanche of rumors regarding the stability of the present price of this metal. Markets have been unsettled from Johannesburg to London from New York and Toronto to Melbourne, despite repeated denials from, American and British authorities of any immediate intention to alter the price. Various economists and financial writers have blossomed forth with predictions that the price would be set at the old level of $20.67 per oz.; some have even been so brash as to state That a price of $15 would be necessary to bring production within control. These effusions, based entirely upon monetary theories and with no regard to production costs, serve to confuse rather than clarify the situation.

Jan 1, 1937

By George Rice

TEN years ago, October, 1914, the author had the privilege of giving an-illustrated address on investigations of coal-dust explosions1 to this Institute at one session of its fall meeting in Pittsburgh, Pa. On the following day, the members had the opportunity of seeing a large-scale explosion test at the Experimental Mine near Bruceton, in which the explosion was made to follow a predetermined path by means of rock-dust barriers of various types, which had prevented the explosive waves from reaching the outlets of, the mine by the shortest paths.2 There was also shown a "home-made rock-dusting machine discharging the dust from a flexible nozzle, and a cement gun operated for the quick erection of a fire or ventilation stopping, such as might be put up after an explosion. Following that paper, there was presented a paper by J. Taffanel, Director of the Lievin testing station in the Pas de Calais, which reviewed the French coal-dust investigations.3 Another important paper was by Howard N. Eavenson, on Coal-mine Explosions Caused by Gas or Dust.4 This paper contains complete lists of explosions that occurred in North America and Great Britain, from early in the 19th century to 1912; in France to 1903, and in Belgium to 1909. No one can look over these lists without being struck by the great number of explosions and the, appalling aggregate of lives lost by them. Although they still happen, we-are hopeful that we have entered on a happier period of freedom from these dreaded visitations, as the result of great increase in knowledge; of the causes and means of prevention. Looking back over coal-dust investigations, there appear to have been periods of advance and of stagnation.

Jan 3, 1925

By Ulick R. Evans

I Should like to commence by saying how much I appreciate the honor which the American Institute of Mining and Metallurgical Engineers has done me in inviting me to visit your country, and to deliver this lecture before you. In choosing a subject, and in deciding what to include and what to omit, I have endeavored to discuss as far as possible work carried out recently in Europe with which you may not be unduly familiar. It would be poor return for your hospitality, and a mistaken sort of compliment indeed, if I were to load the lecture with accounts of researches carried out in your own country—researches with which you are probably better acquainted than I. Not that I have in any way excluded work done on this side of the Atlantic from my survey; one would indeed find it difficult to write on this, or on any other branch of science, without mentioning American work. But I have endeavored as far as possible to concentrate on aspects of the case which will, I hope, be comparatively fresh to you, and I feel fairly certain that this will be in accordance with your own wishes. Films Produced at Elevated Temperatures If a strip of iron is heated strongly at one end, the other end remaining cold, there is produced over the heated portion a layer of oxide scale, the thickness of which diminishes steadily with the distance from the hot end. Somewhere in the middle of the strip, the thickness of the oxidized layer will become comparable to the wave length of light, and over this region the thin films give rise to a series of beautiful interference tints, commonly known as "temper colors." The colors are caused by the fact that two trains of light waves are reflected respectively from the outer and inner surfaces of the oxide film, and the possibility arises that the "crests" of one series of waves may coincide with the "troughs" of the other, thus diminishing the intensity of the light of some particular wave length. The state of affairs in this part of the strip is shown diagrammatically in Fig. 1, in which the film thickness is enormously exaggerated, of course, in comparison with the length of the strip; the

Jan 1, 1929

By Ulick R. Evans

I Should like to commence by saying how much I appreciate the honor which the American Institute of Mining and Metallurgical Engineers has done me in inviting me to visit your country, and to deliver this lecture before you. In choosing a subject, and in deciding what to include and what to omit, I have endeavored to discuss as far as possible work carried out recently in Europe with which you may not be unduly familiar. It would be poor return for your hospitality, and a mistaken sort of compliment indeed, if I were to load the lecture with accounts of researches carried out in your own country—researches with which you are probably better acquainted than I. Not that I have in any way excluded work done on this side of the Atlantic from my survey; one would indeed find it difficult to write on this, or on any other branch of science, without mentioning American work. But I have endeavored as far as possible to concentrate on aspects of the case which will, I hope, be comparatively fresh to you, and I feel fairly certain that this will be in accordance with your own wishes. Films Produced at Elevated Temperatures If a strip of iron is heated strongly at one end, the other end remaining cold, there is produced over the heated portion a layer of oxide scale, the thickness of which diminishes steadily with the distance from the hot end. Somewhere in the middle of the strip, the thickness of the oxidized layer will become comparable to the wave length of light, and over this region the thin films give rise to a series of beautiful interference tints, commonly known as "temper colors." The colors are caused by the fact that two trains of light waves are reflected respectively from the outer and inner surfaces of the oxide film, and the possibility arises that the "crests" of one series of waves may coincide with the "troughs" of the other, thus diminishing the intensity of the light of some particular wave length. The state of affairs in this part of the strip is shown diagrammatically in Fig. 1, in which the film thickness is enormously exaggerated, of course, in comparison with the length of the strip; the

Jan 1, 1929

By Edward Gridley

Members of the Institute who were present at the Amenia, N. Y., meeting, in October, 1877, will remember their visit to the hematite mines, just west of the village of Amenia, and some of them may perhaps recall the deposit of carbonate ore south of the opening made for the hematite. The smelting of this carbonate during the last few months, has produced an iron that seems worthy of being brought to your notice. A little more than a year ago, at the Wassaie Furnace, Dutchess County, N. Y., we made a few hundred tons of iron from a mixture of two-thirds ram carbonate and one-third Chateaugay ore, hoping that it would be suitable for steel purposes, but as the iron showed phosphorus 0.189 per cent., it, was not offered. This iron looked well and seemed quite strong, and gave good results in malleable castings; but no special tests of strength were made. About February 1st, of this year, we began using two-thirds roasted carbonate, and one-third Chateaugay, and noticing that the iron mas stronger than usual, we had two samples tested, which showed tensile strength of 32,014 and 34,176 pounds per square inch. After our stock of Chateaugay ore was exhausted, me put on one-third ram carbonate with the two-thirds of roasted carbonate, and the first test made of the iron, showed 40,008 pounds per square inch. The three tests given above were made by Mr. A. J. Copp and Mr. E. B. Manning, of the Phoenix Furnace, Millerton, N. Y., on a machine of Riehle Bros., Philadelphia. Since these tests were made, they have broken samples made with all carbonate ore as follows: 39,669, 40,816, 41,882, 42,281, 39,902, and 40,130 pounds per square inch. A test taken from the same bed of iron as the last-mentioned (40,130), was broken by Mr. A. Blass, at Irondale Furnace, showing 40,151 pounds per square inch. A sample broken on the Riehle Bros. machine, at Stevens Institute, under direction of Professor R. H. Thurston, showed 40,000 pounds per square inch. Another sample was broken by Professor Thurston on his torsion machine, and gave torsion 7", and tensile strength of 44,500 pounds. And

Jan 1, 1884

By Frederick M. Becket, Russell Franks

WideSpread experimentation has been conducted in recent years to devise a means of preventing intergranular corrosion in austenitic chromium-nickel steels of the 18 per cent chromium-8 per cent nickel type. Additions of titanium to these steels have been used commercially to overcome this dificulty, as advocated by Krupp of Germany, and the latest literature reference concerning improved details of this method is the article by Bain, Aborn and Rutherford.' Extensive work by the present authors has shown columbium also to be a particularly valuable agent for inhibiting intergranular corrosion. This paper does not deal with steels that are largely or wholly austenitic; it is limited to a discussion of the results obtained through the use of titanium and columbium in plain high-chromium steels. For many years titanium has been used commercially by some steelmakers as a cleanser, and to a lesser extent as a hardener, but then in greater percentages. The results of the work to be described disclose that titanium and columbium in proper Proportions act as effective softeners in the high-chromium steels. It is well known that chromium intensifies the hardening effect of carbon in steel; it imparts pronounced air-hardening properties within a rather wide range of chromium content; it provides increased ductility in steels of given strength and hardness; and it enhances resistance to corrosion and oxidation at elevated temperatures to a marked degree. However, in applications requiring a combination of some of these valuable properties great hardness is frequently a detriment rather than a merit. When the objective is a soft and ductile steel having optimurn corrosion resistance, the producer strives for a carbon content as low as commercially practicable. Despite these efforts some of the commercial high-chromium steels low in carbon present considerable difficulties during production, owing to their air-hardening properties, and in consequence the economic result is not favorable. A good example is found in the 5 to 6 per cent chromium steels employed chiefly in the oil-cracking industry. Sleels

Jan 1, 1934

ALABAMA Anniston.-Bretz, J. A. Carrington, F. G. Gerber, A. B. Klugh, B. G. Semple, R. A. Ashland.-Barton, .J. Bessemer.-Abbott, C. E'. Ball, T. L. Beaver, J. J. Ferguson, V. Salmon, H. S. Sehaber, C. F. Whatley, W. J.

Jan 1, 1923

By Bunting S. Crocker

Bunting S. Crocker (Executive Vice President, Kilborn Engineering Ltd., Toronto, Ont., Canada)-This is an interesting report on laboratory-scale testing with pebble charges in 12-in.-diam mills, and the author correctly concludes that useful laboratory testing can be done. In the opening paragraphs he states, "laboratory pebble-milling tests have been considered too difficult to perform and not relevant enough to the actual grinding situation"!! This is definitely not so! The writer published detailed tests done in a 12-in.-diam mill comparing pebble and steel balls in 1952 6 and again in 1954 7. We note that our review paper in 1963 was the only one mentioned in the References. The author states that "a smooth lining is more frequently used in commercial-scale pebble grinding." The writer has designed some 22 plants treating 134,000 tpd in North America and Australia. He has also visited pebble installations in Norway, Sweden, Finland, and South Africa. All these plants correctly use heavier lifters than would be considered standard for ball mills of a similar size. Smooth liners are not recommended in wet ore treating plants. In the 1952 paper6 it was clearly demonstrated in the test work that the individual ore pebbles had to be the same weight-not the same diameter or size-as the steel balls if comparable grinding results were to be obtained. In the tests performed in this study most of the pebbles "passed 50-mm round hole screen openings and remained on a 40-mm round hole sieve. Correspondingly most of the ball mill experiments were performed with 50 or 40-mm steel balls". Thus, the balls and pebbles were essentially the same size or diameter. This means we are comparing apples and oranges. When the feed being ground is fine enough the pebbles will be superior. When the feed is coarse the steel balls will be the better. This is similar to grinding tests in which two sizes of steel balls are compared. In the middle of the article, "Crocker5 has reported that equal weights of pebbles and steel balls achieve the same grinding capacities under similar conditions …", then he proceeds to question this postulate. In the eight reports that I have published, I have never made this statement. What I have said is "for comparable grinding, the individual pebbles in the mill charge should be the same weight as the individual steel balls." This is not the first time that I have been misquoted on this point! It is most unfortunate that this test work has been done with a ball and pebble of the same diameter and not the same weight. It is impossible to properly evaluate the results. I also regret the fact that the screen analyses of the tests were not shown. Many of us who have been 40 years or less in the grinding game can easily read screen analyses at a glance. We can tell whether the top meshes are being ground by the ball charge, etc., etc. The graphs are fine for the summary or to illustrate conclusions but are often misleading in the report itself. However, we do agree that useful pebble grinding tests can be done in a 12-in. mill.

Jan 1, 1973

By Arthur Lake

Within the past few months I have had occasion to visit and examine the area of country in Gunnison county locally called the Gold Belt and extending from the Cebolla river on the west to the head of Taylor Park and the Sawatch range on the east. The region including the "belt" may be hided topographically into two parts, a northern and a southern one. The former is characterized by lofty mountains and deep canons, with occasional parks and wide valleys. Most of it is included in the granitic system of the Sawatch range, with local patches of the Palaeozoic sedimentary rocks, more or less traversed by great bodies of porphyry and other eruptive rocks. The southern portion, on the other hand, is characterized by comparatively low, rounded hills of schist and sahistose gneisses, underlain by coarse, massive granite, which outcrops locally in small, and sometimes in large, patches. These hills also sometimes form table-lands by reason of overflows of lavas, such as andesitic breccia, rhyolite, trachyte and basalt. The leading geological feature is that of a schistose area of width and extent unusual for Colorado. These schists and gneisses are of pre-Cambrian age; but whether assignable to the Algonkian or to some older system has not been determined. The granite which occasionally outcrops from beneath these schists is, I think, of true eruptive character, since its mode of occurrence indicates a once molten, or at least highly viscous or plastic, condition. Thus, we may observe, included in its mass, numerous fragments of schist, almost like a coarse, volcanic breccia. Again, at its immediate contact with the schists, which often stand vertically upon it, the pediments of the latter are fused into the granite, and granite tongues and veins run up amongst the schists, as only a molten or viscous

Jan 1, 1897

By Douglas Bunting

(Wilkes-Barre Meeting, June, 1911.) WITH the gradual exhaustion of the upper veins in the anthracite coal-fields, the problem of mining at greater depths acquires increasing importance and demands the consideration of a number of important factors, one of which is the greater earth-pressure and the consequent necessity of stronger support o for the roof. Under the pillar-and-chamber system, almost exclusively followed in the Northern anthracite-field, the highest economy in mining is generally secured by leaving, on first mining, pillars only sufficient to support safely the overlying strata. As to the necessary size of such pillars, the opinions of mining experts are widely divergent. In establishing the width of chambers and pillars, the thickness of vein and its depth below the surface have received little consideration. For. instance, it has been quite usual to work both overlying and underlying veins with the same width of chambers and pillars, when the lower vein was two and one-half times as thick, and twice as fits below the surface, as the upper. In view of the generally-accepted theory that the crushing-strength of coal-pillars of the same base-area becomes less with increased height, it is probable, in this instance at least, that the most economical mining has not been secured. This question of adequate pillar-support is economically less important down to about 800 ft. than at greater depths; and it is with reference to mining at these greater depths that the study of the subject here presented has been prompted. The necessity of leaving larger pillars, involving greater mining-costs per ton and also smaller yields per acre, is one of the troubles of deep working. To mine without leaving adequate pillar-supports will result, sooner or later, in a squeeze. Limited areas, it is true, have been mined at certain widths of

Sep 1, 1911

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 950°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 I 7 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 2000°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 1/2 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 F. C. Sturges

FOR some time mining engineers have been interested in the possibility of using small diameter shafts, sunk by core drilling, as aids to ventilation and as emergency escapeways. The possibilities are interesting to contemplate, but new methods involve unknowns, and interest tends to remain passive until these unknowns are solved. Recently we have completed two 36-in. diam. core drilled shafts, and wish to make the data on these available to the industry. Both shafts were drilled for the McLain Fire Brick Co., one at its Buckeye mine and the other at its Colonial mine. Both mines are in the fireclay underlying the Lower Kittanning coal. Both mines have been in operation for a number of years [ ] and have required expensive maintenance work on return airways. It was decided that considerable savings could be effected by short-circuiting the air to the surface through small shafts. It was also realized that these holes, located near the present workings, would be of added value as emergency escapeways. Originally, 48-in. diam. shafts were considered, but actual computations showed 36-in. diam. shafts were more than ample at both mines. The shaft at the Buckeye mine is 240 ft deep, and that at the Colonial mine 258 ft deep. Both holes were started in the lower part of the Conemaugh formation and penetrated Conemaugh and Allegheny strata. The holes were started through the unconsolidated materials overlying bedrock by constructing a concrete collar with a wall thickness of 8 in. Then the shafts were drilled by the shot-core drill method. An annular groove was cut in the rock by the abrading action of chilled steel shot run under the shoe of the core barrel. The drilling was done in stages, the core barrel drilling the groove for a certain distance depending on conditions encountered, and then the cores broken off and taken out in sections. The core barrel had a 36-in. o.d. and the finished diameter of the shaft measured almost 37 in. The major questions that could not be answered until the completion of the first shaft were, (I) how much ground water would be encountered, how would it hinder the drilling, and how difficult would it be to seal off, and, (2) how would the walls of the hole stand up. The holes of course could be cased or lined, but that is considerable added expense in any shaft.

Jan 1, 1947