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By W. N. Page

The Red-Ash colliery was the scene, March 6, 1900, of an explosion by which some fifty men were killed or wounded, and much property was injured or destroyed. On March 19th, the earliest date at which entrance was practicable, I examined the mine with the view of determining the extent, cause and point of origin of the explosion, and with the permission of Mr. F. Howald, the manager, and the owners of the property, for whom the examination was made, I present in this paper some results of that examination, comprising the essential portions of my report to them. Location and Plan of the Mine. The Red-Ash colliery, situated on the south side of New river, next below the Rush Run mine, with which the underground workings are connected, is a drift-mine, the entrance to which is on a slope, about 500 ft. above the river. (Fig. 1 is a plan of the workings, in which the points where human bodies were found are designated by numbers.)

Jan 1, 1901

By E. W. Shaw

Continued discussion of the paper of Roswell H. JOHNSON, presented at the New York meeting, February, 1915, and printed in Bulletin No. 98, February, 1915, pp. 221 to 226. See also Bulletin No. 101, May, 1915, pp. 1157 to 1162. E. W. SHAW, Washington, D. C. (communication to the Secretary+). -With the word petrology-literally the science of rocks-being used primarily for igneous rocks, and with geophysics coming to mean the study of lavas and the rocks into which they solidify, contributions to the knowledge of the physical chemistry of sedimentary rocks, even though only ideas which may be used as working hypotheses, become most welcome. The forces which control the formation and subsequent history of the stratified portions of the earth's crust are so little known that no one has yet been able to settle finally the questions as to whether petroleum is of organic or inorganic origin; whether it originated far from its present position, or nearby. Yet both the chemical and physical problems of the rocks in which oil occurs seem to be more simple than those of the igneous rocks. The oil-bearing rocks of western Pennsylvania, for example, have in all probability never been subjected to a pressure exceeding 3,000 or 4,000 lb. to the square inch, nor to a temperature higher than 100° or 200° C. These pressures are considerably below the crushing strength of rocks, and the temperature far below their melting point. The igneous rocks,, on the other hand, have been affected by enormous pressures and very high temperatures and yet steady progress is being made in reproducing in the laboratory the conditions of their formation and subsequent modification within the earth. The solution of the sedimentary-rock problems would seem to be fully as important as those of igneous rocks, and yet we must content ourselves with reasoning backward and our inferences as to what has happened are oftentimes little more than guesses.1 One of the fundamental problems is the origin of the salt water in oil fields. Is the salt water fossil sea water; has it been formed by the solution o-f salts of various kinds from distinct beds, from disseminated par-ticles, and from the weathering of suitable grains in the rock; or is it of deep-seated origin? Its general resemblance to sea water points to the first, but the principles held by some concerning the free circulation of water throughout the earth's crust, and certain peculiarities of composition indicated by the few analyses available, seem out of harmony with such an explanation.

Jan 7, 1915

By C. van Dijk

In Sept., 1960, a steam-drive project was started in the solution-gas drive area of the Schoonebeek field. A part(ern of three five-spots and one four-spot was selected covering an area of 65 acres. The pay in the project area has good lateral continuity and dips slightly to the northeut; it is about SO ft thick and permeability increases from 1,000 and at the bottom to approximately 10,000 md at the top. The oil originally in place was 12.6 X 10' bbl. The oil has an in situ viscosity of about 180 cp. At the start of the steam drive the cumulative primary production due to. solution-ga.7 drive amounted id 4 Percent of the oil originally in place. Reservoir pressure had dropped about 120 psi and no significant primary re-.serves remained. Some 11.3 million bbl of steam (all steam quantities are expressed in barrels of water vaporized) have been injected, resulting in production of an additional 4.1 X I0 9bl of oil, or 33 percent of the oil originally in place. This corresponds to a cumulative oil-stearn rario of 0.37 bbllbbl. It appears that the steam preferentially moves r updip while liquids are produced mainly from downdip wells observations indicate that tile steam flows through only the upper part of the formation. The lateral steam distribution in the pattern is satisfacrory since several prodriction wells hardly reacted and, hence, cori tributcd little to the oil production. Production performance and results from material balance calcutlations agree satisfactorily with the results of large-,scale laboratory experiments. On the basis of these experirmental results the .steam drive, together with a cold water follow-up. is expected to bring ultimate recovery to a value of crt leas: 50 percent of the oil originally in place. No serious production problems have been encountered. However, due to mechanical fuilure, two old prodriction wells and one injection well had to be replaced. An extension of the. steam drive in this area is under connstruction. Introduction The Schoonebeek oil field, discovered in 1943 and developed after World War 11, is situated in the eastern part of the Netherlands. The main oil reservoir in this field is the Valanginian sand. A completely sealing fault divides this reservoir into two areas (Fig. 1): the southwestern part of the sand body where primary production is ob- tained by means of a solution-gas drive, and the remain. der where edge-water drive is the production mechanism. In the greater part of the field the reservoir consists of a single, unconsolidated sand body. The net thickness ranges from 30 to 100 ft and the top is between 2,400 and 2,800 ft below sea level. Formation permeability varies from approximately 10,000 md at the top to values of the order of 1,000 md at the bottom, and porosity is about 30 percent. The reservoir contains a paraffinic oil of 25" API gravity with an in situ viscosity of 160 to 180 cp. Initial oil saturation was high (85 to 90 percent). The relatively large quantity of oil in place (more than 10' bbl), and the low ultimate primary recoveries expected from this field — approximately 15 percent stock-tank oil initially in place (STOIIP) for the water-drive area and 5 percent STOIIP for the solution-gas drive area — clearly indicate ample scope for secondary recovery. Because ies-ervoir and crude characteristics made this field suitable for thermal secondary recovery, a hot-water drive project was started in the water-drive area about 10 years ago. A few years later a steam drive and an in situ combustion project were started in the solution-gas drive area. This paper deals with the performance of the steam-drive project, which was started in Sept., 1960, and which is still in operation. Design of Steam-Drive Project, An experimental investigation of the steam-drive process carried out by schenk in 19561 indicated that under schoonebeek conditions steam injection could be an attractive secondary recovery method. the findings and encouraging results of a pilot test in the Mene Grande field in venezuela,i led to the design of a steam-drive project in the schoonebeek field, Pruject Site and Pattern In 1958 the reservoir pressure in the solution-gas drive area had decreased to about 120 psi, and oil production rates of wells in this area had dropped to 7 to 10 B/D. The cumulative primary production was about 4 percent STOIIP, leaving an oil saturation of approximately 85 percent. In view of the large amount of oil left behind in the reservoir, the solution-gas drive area was selected for the planned steam-drive project. The area in the vicinity of Well S1 3 (Fig. 2) was considered to be suitable since it is at least partly isolated from the rest of the field by faults and the sand is relatively thick (about 80 ft).

Jan 1, 1969

By Frank H. Probert

The mining laws relating to winding-ropes in some of the German provinces are very strict, and severe tests have to be periodically made to see that the ropes in use come up to the prescribed standard. A few notes on the method of testing the ropes and the results obtained may be of interest to the members of the Institute. Permission must be given by the Inspector of Mines to put up a winding-plant; and previous to installation a full report made by the manager, clearly setting forth all details of the shaft, the pit-head, engines and appliances appertaining thereto has to be sent to the mining officials. Accompanying each subsequent report, further details, such as the name of the firm supplying the rope, construction, quality of steel, date of installment, weight per meter, length, particulars of the safetycatches, cages, etc., must be given. At least one spare rope must be kept at ever? shaft ready for use in case of emergency. The rope must hang in the shaft, with cage attached, not less than twenty-four hours before it is used for winding; and it must have undergone a severe test in the manner hereafter described. The laws prescribe that when a rope is continually in use it must be tested every three months; but if the rope is only used every few days it must be tested at least every six months, in the presence of the superior officer; and notification of the date of testing has to be sent to the Bergrath two days before the test is made, in order that he can personally superintend the operation if he chooses. Should winding have been stopped for a long time, it is not necessary to test every six months; but before the rope can be used again it must be tested.

Jan 1, 1901

By D. Torrey

CONSIDERING the two great interests of to-day, in iron upon the one hand and steel upon the other, and recognizing with measurable distinctness the peculiar fitness of each for special services to which they are applied, the proposition to unite the two metals in combination, to bring into simultaneous use the peculiar advantages which each offers, has great attractions. The limited success which has been attained by trying to bring about such combinations of the two metals through the use of fluxes in welding, has led to the general adoption of the opinion that an effective union of the two metals is impracticable, or to the view that if practicable it would have but limited commercial value. Such opinions are, I think, premature and erroneous, and having gained some knowledge of the character and value of the "Wheeler process" for welding iron and steel without the use of fluxes, I submit the following observations and opinions of the matter for your information, with the hope of eliciting such a discussion of the subject as it may require. The principle of the invention is to bring the two metals simultaneously to a welding heat in one pile, and then to manipulate it.

Jan 1, 1879

By John F. Brass

INTRODUCTION The Jim Walter Corporation is a Florida based organization with diverse interests. They have operations in many states and overseas, but the heart of the company is in the southeast. The major part of the corporation's business is home building and everything that is related to it. In 1969, the Jim Walter Corporation gained an interest in coal mining through the acquisition of the U. S. Pipe & Foundry Company which had two slope mines to supply its requirement of metallurgical coal. In the early 1970's it was decided to enlarge the coal mining activities and work was started on the Blue Creek project. This involved sinking four new deep mines (originally there were to have been five but it was later decided that the property could be satisfactorily worked by four mines). They are all located in west central Alabama between Tuscaloosa and Birmingham. They have a combined projected output of 7314 million tons of very high grade metallurgical coal per year to be reached by the mid 1980's. These new mines are all to work the Blue Creek seam which is a very gassy seam and in the area to be mined is found at between 45m (1500 ft.) and 731m (2400 ft.) below the surface. Longwall is the only practicable method of mining to produce the planned outputs in the conditions prevailing and each mine is planned to have three longwalls as the primary production system. Continuous miners will be used just to do the development work for the longwalls. No. 3 Mine is the most advanced of the mines and started the first longwall early in 1979. The second unit started in the middle of 1980 and the third is scheduled for mid 1981. The first longwall at No. 4 Mine is expected to commence production early in 1982. This paper attempts to relate some of the problems that have been encountered and the lessons that have been learned.

Jan 1, 1981

By Robert Goodrich

THERE is a small area about Murray where bedded gold quartz veins occur in Prichard slate. The ore taken from the upper levels, during the early days, was free milling and was treated by stamp milling and amalgamation. There was likewise a considerable placer working. The report is current that tungsten (scheelite) was encountered, which interfered with the amalgamation in the stamp mills and which, in placering, collected in the riffles in the sluice boxes and. clogged them. Consequently, in the quartz mines, the scheelite, when encountered, was rejected, being thrown into the waste, sometimes inside the mine and at other times outside on the clump. The free surface ores became worked out and as the ore in depth became pyritic and rebellious to amalgamation the mines closed down. Last spring tungsten ore brought a high price, near $80 per unit, 60 per cent. W03 and up. This so stimulated the output of tungsten ore that old clumps were explored and waste stored in the mines was gone over. Tungsten was likewise mined from virgin ground. So the district produced quite heavily for a time. The condition now is different. The price is not so high, about $20 per unit W03, which makes profits questionable when mining in new ground. The accumulations from early clay mining have been exhausted and the output has greatly declined. The Golden Chest mine was the heaviest producer of tungsten. The months of June and July were spent at this mine studying the problem. Since the mine had produced both gold and tungsten, we looked forward to the problem of recovering both from what might be called a gold-tungsten ore. A study of the Golden Chest vein showed this not to be so there. While gold ore and tungsten ore come from the same vein, the tungsten ore is confined to a more or less disturbed or distorted portion

Jan 8, 1917

By P. G. Meikle, C. T. Holland

The effect of the condition of friction upon the ultimate unit compressive strength as determined between the compressed faces of precisely machined coal specimens and specially prepared testing platens has been examined. Reductions of 30 to 40% in the ultimate unit compressive strength of these model coal pillars have been verified when lubricants areused between the faces as compared to non-lubricated faces. Increase in unit compressive strength due to confinement at each of three levels of friction appears to follow generally the widely accepted "1/2 power law", i.e. S - C (L/T)1/2 where S is the unit ultimate strength, C is a constant dependent on the coal being tested and the coefficient of friction acting at the compression surfaces and L and T are the least width and thickness of the specimen, respectively. Results suggest that the surface condition of both the specimen and platens should be mentioned in detail in conjunction with compression testing, for although the term (L/T)1/2 is adequate for all practical purposes, the coefficient C is directly dependent on the condition of interfacial friction, all else remaining constant. One phase of the investigation of the confining effect of friction on the ultimate unit compressive strength of small coal specimens has recently been completed at the School of Mines of West Virginia University. The efficient control of ground is of fundamental importance to successful mining operations and the design of support pillars has always presented the mining engineer with a difficult and intriguing engineering problem. Since about the turn of the century, many investigators have performed extensive laboratory tests on small coal specimens in efforts to solve some of the perplexing problems of strata control and pillar action in coal mines. A detailed review of this work would serve no purpose here, but a brief mention of the more significant contributions is in order. Muller16 was perhaps the first to illustrate the effect of confining pressure on the strength of coal specimens in compression. His work has recently been more carefully expanded by the experiments of Murrell17 in Great Britain. Rice and Enzian l9 observed that as the size of cubical test specimens increased, the unit compressive strength decreased. These observations were later supplemented by Lawall and Holland, 13 by Greenwald4,5 et a1 at the U. S. Bureau of Mines, by Gaddy3 as well as by Steart,22 Millard15 et a1 and Evans2 et al. Gaddy3 seems to have been first to quantify this relationship with his equation: S = K/vd , where S - ultimate unit compressive strength (psi), K = a constant dependent on the coal, and D =edge dimension (in inches). Holland7 concluded that the strength of coal as affected by lateral dimensions and thickness could be expressed as where K = a coefficient dependent upon the coal, L = least width of pillar (in.) and T = thickness of pillar (in.). Data supplied by this formula have been compared to the estimated stresses in pillars that had been standing for long periods in coal mines and in the cases so far studied the results indicate that the formula predictions are sufficiently accurate for engineering purposes (Holland11 ). This equation therefore is believed to be the simplest and most accurate tool available at the present time for approximating the strength of a coal pillar (over a specified range of L/T) in situ. However, Holland7,9 has indicated that friction between the coal pillars and the top and bottom rock

Jan 1, 1965

GEORGE S. RICE (written discussion *).-An interesting and important question arose during the coal famine of last winter as to whether the development of new mines should be discouraged on account of the need of labor for the necessary construction and development work at a time of scarcity of labor, more especially as there was an apparent overcapacity of existing mines. The question was referred to the author for his comments, which led to the study of the statistics. The essential results of this study are indicated on the accompanying chart, which illustrates, among other things, the increase in production or consumption, and the estimated maximum production capacity of the bituminous and sub-bituminous mines of the United States. The curves shown are based on data published in the "Mineral Resources" of the United States Geological Survey, and cover the period from 1900 to 1916 inclusive. The following curves have been drawn from the data secured: 1. Average number of days that the mines worked. 2. Yearly total production of bituminous and sub-bituminous mines. 2a. Yearly production of machine-mined coal. 3. Average annual production, equivalent to consumption, for a period of 17 years. 4. Number of coal mine employees each year. 5. Estimated capacity of mines, on basis of 300 working clays in the year. 6. Estimated drop in capacity if no new mines are opened. Considering these curves and the data from which they were constructed: It will be noted that the average number of days the miners worked varied from 193 in 1908 to 234 in 1900, the average being 217. The causes for so many idle days out of the possible total number of working days in the year are well known, namely: (a) Lack of business; (b) lack of cars. Taking the country as a whole, (a) is the important factor, which might otherwise be expressed as being an excess of productive capacity. The yearly production of the mines, which is practically synonymous with consumption, has shown a steady increase, with a few set-backs. An average line has been struck which indicates that the average annual increase of production has been 18 million short tons (16,330,000 metric tons).

Jan 5, 1918

By Basil Prescott

Resume.—The district of Santa Eulalia lies 12 miles to the southeast of the city of Chihuahua, Mexico. The ore deposits occur in a Cretaceous limestone of unknown thickness, overlain by a series of rhyolitic tuffs and flows. The limestone is very gently folded and is cut by several systems of fissures, of which only those with a general north-south strike are pre-mineral and control the deposition. The deposits are of two main types; the earlier silver-bearing iron sulphide type occurs as high-grade tabular bodies resembling veins and as large masses of relatively low-grade ore, The gangue of this type is composed of lime-iron silicates that point toward a close association with an igneous rock mass which has not yet been encountered. In the later silver-lead-zinc type the ore occurs in vertical chimneys and horizontal mantas, principally as oxides, though at a depth of 1,500 ft. some sulphides have been encountered. During oxidation the ore chambers have increased in size owing to the solution of the walls by the sulphuric acid generated in the process, and a substitution has taken place of nearly one-half part by weight of the original constituents of the ore, by newly introduced elements and radicals. Both types are believed to have come from greater depths than those yet reached by mining, and their ultimate source, from a consideration of similar deposits, is believed to be an igneous rock in contact with the limestone at some horizon below. The deposits are believed to have been formed at unusually great depths, as indicated by the proved great vertical range of the silicate minerals coupled with the absence of the igneous rock, and by the practically unchanged character of the silver-lead-zinc ores as far as developed. The volcanic capping is believed to be distinctly later than the economically important ore deposition. Introduction The Santa Eulalia district is situated in the State of Chihuahua, Mexico, and lies about 12 miles southeast of the capital, the city of Chihuahua. (See map, Fig. 1.) The district embraces the Sierra de La

Jan 1, 1915

The next meeting of the Institute, the 113th meeting, will he held in Arizona in the latter part of September, 1916. All papers to be presented at this meeting must be published in the September Bulletin or previously. Manuscripts must, therefore, be in the hands of the Secretary of the Institute not later than July 1. In case manuscripts are given by authors to members of technical committees, or others, they must be in the hands of such persons sufficiently in advance so that they may be actually in the hands of the Secretary of the Institute by the date mentioned. In the past, the Secretary has received urgent requests for special consideration regarding manuscripts that come in after the closing date. In view of the obvious unfairness of this, the Committee on Papers and Publications has instructed the Secretary not to accept for presentation at the Arizona Meeting any papers received later than July 1, 1916.

Jan 4, 1916

By H. R. Page, A. E. Jr Lee

From the inception of zinc retorting on a commercial scale in the United States in 1890,' the retort employed has undergone wide variations in its composition and manufacture, facilitating in part equally remarkable improvements in furnace capacities. The early day hand made clay retort was charged with carbonates or silicates or with coarse dead roasted concentrates mixed with a large proportion of charge fuel resulting in a relatively low zinc burden and fired 24 hr in direct coal fired furnaces. Its modern counterpart is fabricated in hydraulic presses from clay mixtures containing sizeable amounts of either silicon carbide or silica flour, charged with sintered flotation concentrates to more than three times the early day zinc burden and fired 24 to 48 hr in gas fired furnaces. This paper does not attempt to describe in detail the early day clay retort practice as it is well outlined in treatises by Ingalls,2 Lodin,3 Liehig,4 Hofman5 and others. A brief review of clay retort practice is presented together with a description of the major developments since 1912. Clay Retorts The Belgian type retort, both in the circular and elliptical forms, has been used almost exclusively. Typical dimensions of press made clay retorts around 1910 are shown in Table 1. Variations in these dimensions were used at different plants according to local conditions to a maximum inside diameter of 9 in. and inside length of 54 in. However, the effective heat penetration in a 24 hr firing cycle and the tendency of the retort to bend limited the retort size. Use of the elliptical vessel was an attempt to present a stronger cross-section resisting the tendency to bend and to increase the burden without increasing the depth of heat penetration. One exception to the 48-54 in. length was the 60 in. retort used as early as 1905 at Palmerton by means of supporting the last 12 in. at the butt end with a specially designed furnace back-wall. This backwall construction with the 60 in. retort had been developed and used at Bethlehem by G. G. Con-vers and A. B. DeSaulles. An attempt was made at Blende, Colo. to use even larger retorts of the Rhenish type based on European practice and requiring much higher furnace temperatures. Satisfactory plastic clays capable of withstanding these temperatures were not available, and the plant never operated successfully. PREPARATION OF BATCH Composition of the clay retort by weight was 40 to 50 pct raw clay and the balance "grog." Generally speaking the mix consisted of 7 parts plastic clay to 9 parts grog by volume. Principal source of the clay used was the Cheltenham vein—sometimes referred to as "St. Louis city clay." A typical analysis was A12O3-31.0 pct, SiO2-50.0 pct, Fe2O3-2.5 pct, MgO-0.3 pct, CaO-1.5 pct and loss on ignition 14.0 pct. At the smelter the clay was weathered whenever possible and then crushed to 0.08 in. or finer. Grog consisted of calcined adobies, cleaned retort scrap and cleaned refuse fire brick such as old furnace brick, blast furnace linings, and others. Saggers from ceramic plants and calcined flint clay were later used. The grog materials were ground to 0.12 m. or finer. Occasionally coke dust up to 10 pct of the mix was substituted for a part of the grog following European practice.² Particle size of the grog has a major influence on the retort properties—the larger the grain, the better can the retort withstand thermal shocks, resist bending at furnace temperatures and resist corrosion from slag; the smaller the grain, the lower the loss of zinc vapor through the retort walls. Grog forms the skeleton of the retort, and the clay shrinks around its grains to act as a binder. In the drying process, the grog has a stabilizing effect on the drying rate, decreasing shrinkage and giving up previously absorbed water to the surrounding clay to minimize the danger of cracking or checking.² Grog and clay were mixed through a horizontal pug mill with 10 to 20 pct water added, depending on whether the retort was to be formed by hand or mechanically, more water being required for the hand process. The batch or "mud" extruded from the pug mill was cut in convenient lengths for handling, stacked in piles or in special rooms, covered with wet burlap and allowed to "rot" or age from 1 to 8 weeks to increase plasticity. HAND MOLDING If the retort was to be molded by hand, the mud was repugged after the rotting period and given to the molders. Their molds consisted of 3 sheet iron or wood cylinders, each one third the retort length and defining the outer shape of the retort. Beginning with the bottom section, mud was placed in the form and tamped with a rammer

Jan 1, 1950

By Klaus-Peter M. Hanke

INTRODUCTION In early 1979 Jim Walter Resources, Inc. (JWR) of Brookwood, Alabama approached TMCI Construction, Inc (TMCI) to make a proposal on a program that involved the sinking of up to 10 ventilation shafts of approximately 6.7 m (22 ft) diameter and ranging in depth from 500 to 700 m (1650 to 2300 ft) for the JWR coal mines in Alabama. At this time TMCI was already constructing the first spiral underground bunker (capacity 2000 tons) in North America for the JWR organization at their No. 4 mine in Alabama. The TMCI proposal was based on the use of the mos modern large diameter shaft boring machine rather than sinking the shafts using the conventional drill blast-muck technique. The proposal was made based o: the experiences by the parent company, Thyssen Schachtbau, which has been using this type of machin in Germany for shaft boring since 1971. As a result of the TMCI proposal JWR issued a purchase order to TMCI for the construction of four 6.7 m (22 ft) diameter, concrete lined, unfurnished ventilation shafts ranging in depth from 500 to 700 (1650 to 2300 ft). An order was thus placed with WIRTH Machinen- and Bohrgeraete Fabrik GmbH, in Germany for the manufacture of a model 650/850 E/Sch "Schachtbohrmaschine" (Vertical Shaft Borer = V-mole which arrived on site in Alabama in early 1981. The first V-mole GSB 450/500 was introduced in Germany in 1971 and was capable of enlarging in one step a pre-drilled 1.2 m (4 ft) pilot hole to 4.5 - 5.0 m (14.7 - 16.4 ft). This machine has sunk 9 staple shafts and deepened one surface shaft for a total of 2360 m (7740 ft) of shafts. On the last shaft boring operation in 1978 the machine was converted as an experiment to drill without a pilot hol using a hydraulic pumping system to remove the cutting debris. A second generation machine, the SB VI 500/650, was introduced in 1977 for enlarging the pilot hole to a range of 5.0 - 6.5 m (16.4 - 21.3 ft) diameter. This machine is still in operation and has already drilled well over 2000 m (6500 ft) of shaft. The third generation of V-mole, the SB VII 650/85( for diameters from 6.5 to 8.5 m (21.3 to 27.9 ft) was: commissioned in May 1980 and has been used for two surface shaft deepenings totalling 606 m (1990 ft) with another scheduled for 1982. The main advantages favouring the use of such V-moles were identified as: 1) A reduction in manpower to the crew required in a conventional shaft sinking operation. 2) A considerable reduction in time to complete a shaft compared to conventional techniques. 3) The use of the V-mole eliminates many of the hazards encountered in conventional sinking. Based on the successful performance of the first three V-moles in Germany, Thyssen Schachtbau decided to employ this principle abroad. In 1980 a second machine of the third generation was built and is now operated by TMCI Construction, Inc. in Alabama. The first shaft was completed at the end of 1981 and this paper describes the method of operation including some unique aspects not attempted on prior V-mole operations and some of the statistics arising out of the experiences during the first shaft boring operation. THE NO. 7 MINE FAN SHAFT SITE Jim Walter Resources, Inc. was formed in 1970 to exploit the coal field in Alabama on the southern tip of the Appalachian coal field. The coal reserves amount to around 650 million tons of mainly good quality coking coal of which about 350 million tons are to be extracted over the next 30 years. Shaft sinking and preparatory work began in 1972, and at present 6 mines are producing around 5.4 million t.p.a. Annual production is to expand to 10 million t.p.a. as soon as possible, and the ventilation shafts to be sunk by TMCI play a vital role towards attaining this goal. The first shaft site is located at the No.7 mine, near Brookwood, Alabama. The actual location of the shaft relative to the production shafts is shown on the mine plan (Fig. 1), which also shows the room-and-pillar extraction system used at present. The mine plan further shows the conveyor route used for the muck removal. The geological survey showed that the strata consisted of horizontal layers of mainly sandstone, sandy shale and shale interspersed with several coal seams. The seam being extracted at the No. 7 mine is a combined seam made up of the Blue Creek and Mary Lee seams at a depth of 513 m (1682 ft) and having an average seam thickness of about 2 m (6 ft). At the beginning of September 1980 the surface site preparation and pre-grouting work was completed by JWR, and TMCI was able to commence with Stage I of the shaft sinking program - the drilling of the pilot hole.

Jan 1, 1982

CONTENTS PAGE WARNER, ROBERT K.-Efficiency of Screening. Discussed by Robert C. Canby,Robert K. Warner, W. 0. Borcherdt, V. E. Flanagan 1 FAHRENWALD, A. W.-Surface Reactions in Flotation. Discussed by Arthur F. Taggart, A. M. Gaudin 7 Efficiency of Screening Discussion of the paper of ROBERT K. WARNER, presented at the New York Meeting, February, 1924, and issued, as Paper No. 1293-M, with MINING AND METALLURGY, January, 1924. ROBERT C. CANBY, Wallingford, Conn.-On page 2, the author speaks of the importance of using a testing sieve clothed with the same material as the screen being studied; does he regard this as essential? In practice, how does the screen efficiency calculated by the difficult-grain method compare with those computed by the usual undersize method? Is it possible by the author's method, to obtain a measure of efficiency that can safely be compared with that of another screen sizing a different ore? ROBERT K. WARNER.-The use of a testing sieve clothed like the mill screen being studied is desirable but not essential for routine work about a mill where the character of the feed varies little, as regards the shape of the grains and the sizing test. Close agreement in size of aperture between the sieve and the screen is, however, important if the data obtained are to be valuable to others working on a different feed. If the testing sieve is coarser than the mill screen, the efficiency will appear unfairly low; whereas if it is too fine one might get efficiencies greater than 100 per cent. because of the apparent oversize in the undersize. In either of these cases the usefulness of the difficult-grain method of calculation would be reduced. As screens wear, the undersize grows appreciably coarser so that it is desirable to test screens only when the cloth is new or else to provide testing sieves that are worn about the same amount. The futility of making efficiency observations on a. distorted or broken screen surface is obvious.

Jan 6, 1924

By Herman Garlichs

The development of the extensive Southeast Missouri deposits greatly preceded that of the Iowa and Wisconsin deposits. It began about 1720 at Mine La Motte and other localities, and has continued uninterruptedly to the present time. It is estimated that this district produced about 184,000 tons of lead in the year 1916, having a value of approximately $25,000,000. This is a new high record. The Southwest Missouri orebodies were hardly known before 1845 and were not extensively developed until 1870. Lead production in Southwest Missouri for 1916 was 38,788 tons. Oklahoma produced 14,399 tons and Kansas 2,345 tons, making a total for Southwest Missouri, and the adjoining districts in Oklahoma and Kansas, of 55,532 tons of lead. The first metallurgists in Missouri were the Indian and the hunter. The early settler learned to procure his bullets either by melting ore in his camp fire or by throwing pieces of galena on an old stump and depending on the usual roasting and reaction method to obtain the metal. A little more refined method was to arrange two flat stones in the form of a, V. Wood and galena were placed in the furnace and ignited.. As heat developed, some of the metal was extracted and molded into bullets. The next step in utilizing the lead ores was the log hearth. All lead in Missouri was smelted in this crude affair before 1820. It was built on sloping ground, and consisted of a hearth of stone, surrounded on the front and two sides by a stone wall. The wall was 7 it. high in front. The top and rear end were left open and in front an arch or opening was made, forming the eye of the furnace. In front of this a pit was dug in the ground, to receive the molten metal. Large logs were rolled in at the back and made to rest upon ledges formed inside, to raise them from the hearth and to give a draft. These logs filled the entire width of the furnace. Small split logs were then set up around the two sides and the front, and the ore was piled on until the furnace was full. Finally, the mass was covered with logs and fuel until the ore was completely surrounded. A gentle heat was started, which was raised very gradually. After 12 hr. the heat was increased and continued for 12 hr. more, 24 hr. being required for each smelting. The furnace treated about 5,000 lb. per 24 hr. and the ore yielded about 50 per cent. of its metallic lead. A considerable quantity of the ore was not desulphurized, and fell between

Jan 1, 1918

By T. S. Spice, H. L. Lovell, R. W. Utley

With the increased demand for lightweight aggregate, such materials have been manufactured from slags, clays, slates and, to a minor extent, the refuse of coal preparation processes. The latter source is attractive due to current interest in eliminating air pollution and improving the unsightly appearance of refuse piles as well as potential economic advantages, if a satisfactory product can be manufactured. A study has been made relating the composition of coal refuse, expressed as washability data, and the properties of a resulting aggregate. Three different refuse samples (bituminous and anthracite) were separated at gravities ranging from 1.4 to 2.96. Their chemical composition was determined. Aggregates were prepared in the laboratory from these fractions before and after ashing, at temperatures between 1250' and 1350°C. The specific gravity, expansion and compressive strength of the products were measured. The response of the refuse to bloat is related to the chemical composition of the ash to yield a glass of desirable viscosity and to the presence of a material which will form a gas in the appropriate temperature range. Carbonaceous material within the particle is a negative factor since after the glassy phase forms, oxygen cannot enter the mass thus inhibiting bloating. Precombustion of the refuse is also detrimental since it removes substances of gas-forming capability. The fraction sinking at 2.60 yielded an aggregate which had the lowest specific gravity, greatest expansion and greatest compressive strength. The purpose of this study was to seek and evaluate possible relationships between the composition of coal refuse, as expressed by washability data and factors affecting the production of a lightweight aggregate from refuse by bloating. The study was related to an intense program in the Department of Mineral Preparation at the Pennsyl- vania State University to study the possible elimination of air pollution from coal refuse fires by producing a low-carbon secondary refuse. Another phase of this program is considering possible uses of such a secondary refuse. Thus one study was to determine if a lightweight aggregate could be produced from this new refuse material and what effect the nature of refuse beneficiation would have upon the characteristics of a commercial aggregate product. It was hoped that these studies would assist in the design of a process and in establishing the optimum specific gravities at which refuse separation could be made in respect to the production of a low grade fuel from the float product and a raw material for lightweight aggregate from the sink product. The combination of producing a saleable low grade fuel and a lightweight aggregate, while at the same time minimizing disposal costs and air pollution contributions would be attractive to the coal mining industry. MECHANISM OF BLOATING Austin, Nunes and Sullivan,1 Ehlers,3 Everhart, 4 Riley7 and others, have stated that bloating is an expansion process in which two essential conditions exist simultaneously. These are: the development of a glassy phase over a relatively wide temperature range (the viscosity of this glassy phase plays a major role in the retention of the evolved gases) and the evolution of gases (SO2, SO3, CO2 and O2) from mineral decomposition in sufficient quantities at appropriate temperatures to form a cellular structure in the mass. Unless these two "type" reactions take place at the same time or in overlapping ranges of time and temperature, cellulation will not occur. In this event, the gas will escape or the melt will be too fluid to retain it; or if retained, the cells may be too large or the walls too thin for good strength in the product. The stages that occur when clays and shales are heated in an oxidizing atmosphere may be summarized as follows: 1) Drying and removal of free water to 200°C. 2) Absorbed water dehydration period (200 to 480°C). 3) Chemically-combined water dehydration period (480 to 700°C).

Jan 1, 1965

By M. Hansen, C. C. Wang

THE 6 phase of the Cu-Sn system (Cu,Sn, = 32.53 pct Sn) was considered to be stable down to room temperature until Owen and Iball showed by X-ray analysis that it undergoes a eutectoid decomposition into the a and e phases at a temperature near 300°C. Their findings have been confirmed by other investigators,2-5 a11 of whom used X-ray diffraction studies of powdered specimens. The transformation temperature has been reported by various authors as lying between 300" and 380°C. To our knowledge, no microscopic evidence of the eutectoid reaction has been published, as yet. It was in connection with other work that longtime annealings at 300°C were conducted with alloys containing 22, 32.6, and 35 pct Sn (nominal compositions) prepared from electrolytic copper and electrolytic tin and cast in a small iron mold. The specimens were first annealed at 700°C for 5½ hr and quenched from 400 °C after cooling to, and annealing at, that temperature for 9½ hr. After this treatment the sample with 22 pct Sn was cold worked for 15 pct reduction in thickness; the 32.6 and 35 pct alloys were too brittle to be cold worked. The final anneal at 300°C was done in a sealed pyrex glass bulb for 63 days. Fig. 1 represents the microstructure of the annealed alloy with 22 pct Sn. It shows primary a solid-solution crystals in a matrix of the (a + 8)-eutectoid formed at 520 °C. The a crystals contain 8 inclusions which precipitated from a due to the decrease in the solubility of tin in copper with fall in temperature. The decomposition of the 8 phase into a (light) and E (dark) can clearly be seen in some areas of the (a + 8)-eutectoid matrix. However, the decomposition is far from being completed, even after the long-time anneal used. Figs. 2 and 3 depict the microstructures of the annealed 32.6 pct alloy, which was intended to consist entirely of 6 phase. Due to some loss in copper on melting, the composition is slightly higher in tin. Accordingly, the sample after quenching from 400 "C consisted of grains of the 8 phase with the e phase arranged along the grain boundaries and as a Wid-manstatten precipitate within the 6 grains. After an annealing time of 63 days, the eutectoid decomposition, 6 + a + e, is still in its early, initial state. The pearlitic type of eutectoid is nucleated at the c crystals, preferably those surrounding the grain boundaries of the 8 phase. The appearance of the (a + €)-eutectoid in the 35 pct Sn alloy was very similar to that shown in Figs. 2 and 3. Again, the eutectoid decomposition nucleated at the 6 phase and proceeded from there into the 8 areas. As is evident from Fig. 2, the rate of decomposition of the 6 phase at 300°C is extremely sluggish: In the representative area shown only about 5 pct of the 8 phase has undergone the eutectoid trans- formation after annealing for 63 days. On the other hand, Owen and Williams2 have reported that X-ray photographs of powdered samples with between 17.5 and 34 pct Sn did not show 8 lines after annealing for only five to seven days at 300°C. This is in accordance with the fact that the rate of reaction in powdered samples is higher than in massive specimens, as a result of the excess free energy of the powder (due to its greater surface area). The authors wish to express their thanks to C. A. Johnson for preparation of the micrographs.

Jan 1, 1952

By R. S. Archibald

The subject of this paper is particularly appropriate, since Daniel C. Jackling was active in early exploitation of taconites on the Mesabi Range and contributed greatly to their later economic development. After his success in developing the low grade porphyries in the West he looked for more worlds to conquer, and so undertook the concentration of low grade taconites, from which he hoped to produce commercial ore. Mesabi Iron Co., organized in 1919, actually did produce merchantable sinter from taconites in 1922. The enterprise did not prove commercial at the time, as cost of production was high and the surplus of iron ore in the Lake Superior District already contributed to a low market price, but this same property has since been developed by Reserve Mining Co. The work done by Jackling had much to do with the success of this later venture. Iron ore was first discovered on the Marquette Range in Michigan in 1845. The first ore from this range came from the old Jackson mine in Negaunee in 1848. During these early years this property and others on the Marquette produced limited tonnages for transportation to Lower Lakes and for consumption by charcoal furnaces in the area, where a plentiful supply of hardwoods furnished the charcoal for iron smelting. The Marquette and other ranges tributary to Lake Superior did not produce much ore for transportation to Lower Lakes until the locks at Sault Ste. Marie were constructed, but from the time the first lock at the "Soo" was completed in 1855 these ores were transported through the locks to consumer centers in the East. The Marquette Range has a history of continuous production from 1848 to the present day. In 1877 the Menominee Range was discovered at the location of the old Penn Mines near Vulcan, Mich. There were further discoveries of ore on the old Menominee Range and at Florence, Wis., in 1880 and at Iron River and Crystal Falls, Mich., in 1882. Penn Mines continued in operation from the date of their discovery until they were finally exhausted by Pickands-Mather in 1945, but operations still continue on the Menominee in the Crystal Falls and Iron River districts. Ore is shipped to a lake port at Escanaba, Mich.

Jan 1, 1960

By Jennings S. Cox

(Glen Summit Meeting, June, 1911.) THE following notes, prepared in 1908, as the result of a personal examination and extensive explorations under my direction in 1906, have been revised and greatly augmented after two subsequent visits and further explorations in 191.0. The hard iron-ores of the south coast of Oriente Province, in the island of Cuba, have been known for many years, in fact one mine has been in operation for more than 20 years ; but it is only within the past four years that the large mantle-deposits of brown ores of the north coast have attracted serious attention. A. C. Spencer 1 has described these ores in a paper, entitled Three Deposits of Iron Ore in Cuba, which refers to the Mayari, Moa, and Cubitas deposits. C. M. Weld,2 in a scholarly paper on The Residual Brown Iron-Ores of Cuba, has discussed the character and possible genesis of these deposits with particular reference to the Moa field, calling attention also to the Taco and Navas fields in addition to those described by Spencer. Descriptive articles with reference to the Mayari field, and a few words about the Moa deposit, have also appeared.3 It is the purpose of the present paper to give some account of the exploration of the Moa deposit and to call attention to its commercial importance. LOCATION. The Moa district is situated on the north coast of the . Province of Oriente, which is the most easterly province of the island of Cuba. It lies on the northern or seaward slope of 1 Bulletin No. 340, U. S. Geological Survey, pp. 318 to 329 (1908). 2 Trans., xl., 299 to 312 (1910). 3 Iron Age, vol. lxxx., No. 7, pp. 421 to 426 (Aug. 15, 1907), and vol. lxxxi., No. 15, pp. 1149 to 1157 (April 9, 1908).

Mar 1, 1911

By F. J. Fayers, R. L. Perrine

For many years the problem of increasing ultimate recovery of oil from a reservoir has been a subject of interest to the oil industry. At present, a standard secondary recovery technique is to flood the reservoir with water once natural forces of the system are depleted. The permeability characteristics of the rock, together with interfacial forces which exist between the two fluids and the porous material. allow only part of the oil in the rock formation to be recovered by the water-flooding process. Laboratory experiments have shown that when detergents are added to the flooding water, the flow properties of oil in a porous rock arc improved. This improvement results in greater oil recovery. Even though higher recoveries are obtained using detergents, the extra cost associated with the method may make it economically unsatisfactory. To establish economic feasibility requires experimental laboratory measurement of fluid and rock properties together with a calculational procedure to predict field results. Pilot tests in the field may also be required. This paper presents a mathematical procedure for predicting detergent flood behavior. A number of reasonable simplifying assumptions are made to obtain a tractable system of equations. The resulting method may be regarded as a generalization of Buckley-Lev-crett' theory to solve simultaneously for water saturation and detergent concentration profiles. Consequently this type of calculation can be used in much the same manner as Buckley-Leverett theory is used to evaluate a water flood. The mathematical method employed is the so-called method of characteristics for solving hyperbolic partial differential equations. This application of the method is instructive because several other problems in reservoir analysis may be analyzed in a similar manner. FLUID FLOW EQUATIONS The flow behavior of oil and water without detergent through porous reservoir rock is usually expressed by Darcy's law and a continuity equation for each phase. If gravitational forces are neglected, vrg = kr (s)µr v p rg.........(1) vrg = kr (s)µr v p rg.........(1) ? pa va = F ?/?t (sd) where v is the volumetric flux, k represents permeability, S is saturation, and P the pressure. Fluid viscosity is denoted by 11, p is the density, 4 is porosity, and I the time. The subscripts w and o refer to water and oil. The capillary forces in the system lead to a difference in pressure between phases. This may be expressed by the relation: Pm -P0 =Pc(S).........(5) The capillary pressure? P,, is determined experimentally as a function of saturation. DETERGENT BALANCE EQUATION Addition of detergent to the water phase affects the flow properties of the system. Capillary pressure and relative permeability become functions of detergent concentration, c, as well as fluid saturation. That is: P--P (S,c) and k,,,.- - -- k,,,. At low oil saturation values, the presence of detergent reduces capillary pressure and reduces the relative permeability ratio. Detergents are surface-active and adsorb on rock surfaces. The quantity adsorbed, a, will depend on the concentration c of detergent solution and may also depend on the saturation; thus, a = a(S,c). In this analysis we have assumed that the detergents used are water soluble but not oil soluble, and their movement is therefore restricted to the water phase. The dispersion of detergent in water is neglected by comparison with transfer by flow within the system. Detergent movement is determined by the water flux and adsorption: ? cv m = -F ?/?t (cs + a) We will assume that the quantity of detergent in solution is too small to significantly change water density or viscosity. When the permeability and capillary pres-