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By Shiou-Chuan Sun

THIS paper presents the design and operation of a frothmeter capable of measuring the frothing characteristics of pine oils and other frothing reagents. The experimental data show that the froth-ability of pine oil is governed by: 1—rate of aeration, 2—time of aeration, 3—height of liquid column, 4—chemical composition of pine oil, 5—pH value of solution, 6—temperature of solution, and 7—concentration of pine oil in solution. The effect of mineral particles on the behavior of froth also was studied, and the results can be found in a separate paper.' The results also show that the relative froth-abilities of pine oils in the frothmeter generally correlate with those in actual flotation, provided that other factors are kept constant. In addition to pine oils, the other well-established flotation frothers were tested, and the results are included. In this paper, compressed air frothing is the frothing process performed by means of purified compressed air, whereas sucked air frothing is the frothing process accomplished by purified air sucked into the glass cylinder by a vacuum system. The term vacuum frothing denotes that froth was formed by degassing of the air-saturated liquid under a closed vacuum system. Apparatus The frothmeter, shown in Fig. 1, is capable of re-producibly measuring the volume and persistence of froth as well as the volume of air bubbles entrapped in the liquid and is capable of being used for compressed air frothing, sucked air frothing, and vacuum frothing. Fig. la shows that for compressed air frothing, the apparatus consists of an airflow regulating system, 1-3; a purifying and drying system, 4-8; a standardized flowmeter to measure the rate of airflow from zero to 500 cc per sec, 9; and a graduated glass cylinder, 13; equipped with an air regulating stopcock, 10; an air chamber, 11; and a fritted glass disk to produce froth, 12. The fritted glass disk, 5 cm in diam and 0.3 cm thick, has an average pore diameter of 85 to 145 microns. The pyrex glass cylinder has a uniform ID of 5.588 cm and an effective height of 63 cm. The inside cross-sectional area of the glass cylinder was calculated to be 24.53 sq cm, or 3.8 sq in. For sucked air frothing, Fig. lb shows that the apparatus for compressed air frothing is used again, with the following modifications: 1—compressed air and its regulating system, 1-3, are eliminated; and 2—a vacuum system, 16, equipped with a vapor trap, 15, and a vacuum manometer, 17, is added. The vacuum system can be either a water aspirator or a laboratory vacuum pump. Any desired rate of airflow can be drawn into the glass cylinder, 13, by adjusting the opening of the air regulating stopcock, 10. The sucked air stream is cleaned by the purifying and drying system, 4-8, before entering the glass cylinder, 13. When this setup is used for vacuum frothing, the air regulating stopcock is closed. The frothmeter has been used for almost 3 years and has proved to give reproducible results, as illustrated in Table I. With a magnifying glass and suitable illumination, the frothmeter also can be used to study the attachment of air bubbles to coarse mineral particles.' Experimental Procedures Except where otherwise stated, the data presented were established by means of the compressed air method. The volume and persistence of froth were recorded respectively at the end of 4 and 6 min of aeration at a constant rate of airflow of 29.3 cc per sec, which is equivalent to 71.6 cc per sq cm per min, or 462.6 cc per sq in. per min. The aqueous solution for each test, containing 1000 cc of distilled water and 19.2 ± 0.5 mg frothing reagent, was adjusted to a pH of 6.9 0.2. The volume of froth is expressed as cubic centimeter per square centimeter and is equivalent to the height of the froth column (the distance between the bottom and the meniscus of the froth). The volume of froth was obtained by multiplying the height of froth by the cross-sectional area of the glass cylinder, 24.53 sq cm. Before each test, the glass cylinder, 13, was cleaned thoroughly with jets of tap water, ethyl alcohol, tap water, cleaning solution, tap water, and finally distilled water. The cylinder with stopcock,

Jan 1, 1953

By Shiou-Chuan Sun

DISCUSSION shiou-chuan sun (author's correction) - Too late for revision, the author realized that the data on stability of froth presented in this paper were not based on the same height of froth column. It is recommended herein that the stability of froth is better calculated by dividing the time elapsed until the froth disappeared by the volume (in cu cm per sq cm) or height of the froth column, and is expressed in sec per cm. The stability data in Tables I, II, and III of this paper were recalculated, in accordance with this recommendation, and the results are shown respectively in the accompanying Tables VII, VIII, and IX. Compared with Tables I, II, and III of the paper, the stability data of Tables VII, VIII, and IX are more reliable in disclosing the persistence of froth of various frothers having different frothabilities. For example, the conclusion given in the paper that the stability of froth is generally proportional to the volume of froth becomes invalid when the stability of froth is expressed in sec per cm, as shown in Tables VII, VIII, and IX. Table VII shows that both the stability and volume of froth increase with the addition of small amounts of n-hexyl alcohol, and that further additions of n-hexyl alcohol decrease the stability but increase the volume of froth. Table VIII indicates that the stability of different types of frothers is governed by their chemical composition and is difficult to predict from their frothability. Table IX shows that the stability of different frothers with similar nonpolar groups tends to increase with a decrease of the hydro- philic property of their polar groups, and consequently with the decrease of their solubilities. This is further substantiated by the fact that the stability of frothers, as listed in Table IX, is generally lowered by the solvent effect of absolute ethyl alcohol. Table IX shows also that the stability of froth tends to increase with the decrease of the volume of froth* provided the frothers used have similar nonpolar groups and dissimilar polar groups.

Jan 1, 1953

By Shiou-Chuan Sun

DISCUSSION shiou-chuan sun (author's correction) - Too late for revision, the author realized that the data on stability of froth presented in this paper were not based on the same height of froth column. It is recommended herein that the stability of froth is better calculated by dividing the time elapsed until the froth disappeared by the volume (in cu cm per sq cm) or height of the froth column, and is expressed in sec per cm. The stability data in Tables I, II, and III of this paper were recalculated, in accordance with this recommendation, and the results are shown respectively in the accompanying Tables VII, VIII, and IX. Compared with Tables I, II, and III of the paper, the stability data of Tables VII, VIII, and IX are more reliable in disclosing the persistence of froth of various frothers having different frothabilities. For example, the conclusion given in the paper that the stability of froth is generally proportional to the volume of froth becomes invalid when the stability of froth is expressed in sec per cm, as shown in Tables VII, VIII, and IX. Table VII shows that both the stability and volume of froth increase with the addition of small amounts of n-hexyl alcohol, and that further additions of n-hexyl alcohol decrease the stability but increase the volume of froth. Table VIII indicates that the stability of different types of frothers is governed by their chemical composition and is difficult to predict from their frothability. Table IX shows that the stability of different frothers with similar nonpolar groups tends to increase with a decrease of the hydro- philic property of their polar groups, and consequently with the decrease of their solubilities. This is further substantiated by the fact that the stability of frothers, as listed in Table IX, is generally lowered by the solvent effect of absolute ethyl alcohol. Table IX shows also that the stability of froth tends to increase with the decrease of the volume of froth* provided the frothers used have similar nonpolar groups and dissimilar polar groups.

Jan 1, 1953

By A. Brown

RAPIDLY rising world demand for mineral products has accelerated depletion of the more readily accessible ores, particularly those of premium grade. Operations must proceed at a faster rate to deeper horizons, where there are problems of ground control that restrict economic recovery of the mineral and introduce new hazards to mine personnel. For some of the new Canadian mining fields severe problems of ground stress are still in the future, but much of Canada's mineral output comes from well established mines that have reached considerable depths. It must also be anticipated that the depletion of premium grade ores will eventually necessitate mining low grade material at competitive cost, and unsolved problems of ground control can materially increase costs. Studies of ground stress phenomena in underground workings are therefore considered a present-day necessity, especially since these investigations must be a long-term effort owing to the highly complex nature of the problem and the difficulty of making full-scale observations. In recent years the Mines Branch of the Department of Mines and Technical Surveys in Canada has conducted such an investigation. This report of the work is primarily concerned with stress problems in some Canadian coal mines and, in particular, with mine bumps. Mines Branch investigations of this and related problems are carried out by a small team of mining engineers and physicists who, with assistance from geologists, pool their special skills in an integrated program. These Canadian investigations combine a field and laboratory approach. Since data obtained under contrasting mining conditions allow useful comparisons of ground behavior under stress, this article will present data gathered under a wide range of Canadian coal mining conditions. Two contrasting longwalls in eastern Canada will be discussed and two dissimilar pillar extraction operations in western Canada. Locale of Observations In eastern Canada observations are being made in the collieries at Springhill, Nova Scotia. The district's No. 2 mine, which has a history of bump occurrences dating back to 1917, is currently being worked at a depth of 4200 ft. The mining method is longwall retreating along the strike and the practice is to retreat three walls in line, each about 350 ft long. There is no machine cutting or shooting—the coal is hand-picked and hand-loaded onto face conveyors. There are three rows of hardwood chock block supports parallel to the coal face. These supports are built during the two daily coal loading shifts, and the rear line is withdrawn during the midnight shift. The goaves are partially packed by stone midwalls 12 ft wide which are built forward daily in step with the face advance. The seam dips at 15° and averages 9 ft of strong, well cleated bituminous coal enclosed in competent beds of shales, sandy shales, and sandstones of Carboniferous age. The bumps rarely disturb the retreating longwall faces but manifest themselves as violent upthrusts of pavement in the levels leading to the walls. Observations have also been made in Springhill's No. 6 mine, which operated in a 5-ft seam of coal at a depth of about 2500 ft. Although No. 6 mine was

Jan 1, 1959

By E. C. Olsen

This paper describes the unusual hazardous conditions encountered in mining thick, inclined coal beds under heavy cover in Carbon County, Utah. This description includes heavy roof conditions, sloughing ribs, steep pitches, and severe bouncing experience. Emphasis is placed on the top bench system of mining and development in the top half of a 16-ft seam, under the roof rock, which is bolted, and the recovery of the bottom coal on retreat; success of bolted roof and ribs; and shuttle car haulage on pitches up to 25%. The use of protective cages for operators of face equipment, the guarding of belt conveyor systems to prevent injury associated with deliberate entry, and the positive lockout of electrical equipment is described and illustrated. Most coal mining areas of the Western United States are characterized either by thick beds, steep pitches or heavy cover. Individually, each of these may present inherent safety hazards that in- fluence mining. When all are met at one location, the combined effect is to make the safe operation of the mining property a challenge to management. These combined conditions are met in the coal mines of the Columbia-Geneva Steel Div. of the United States Steel Corp., in southeastern Utah. This coal bed measures up to 16 ft in thickness, the pitch is as much as 25% and mining is under a cover approaching 3000 ft. The Sunnyside coal bed is found in the Book Cliff Mountains, a great escarpment extending from west of Castle Gate, Utah, some 180 miles to Grand Junction, Colo. West of Green River, Utah, the Book Cliffs swing sharply to the north because of the domelike uplift of the San Rafael Swell to the southwest. It is in this northerly extension of the Book Cliffs that the Sunnyside coal in the Blackhawk Formation of the Mesa Verde Group, Upper Cretaceous Age is found. The strata are transected by numerous faults with from a few inches to as much as a hundred feet of displacement. All of these faults have had some effect on the mining operation, either in roof control or, in the case of major faulting, the pattern of mine development. MINING PRACTICES Mining has generally been confined to the Lower Sunnyside Seam which measures from 7 to 16 ft in

Jan 1, 1963

By Mehmet Guney

At the Armatcuk mine in Turkey, various modifications of sublevel methods have been tried during the last decade in an effort to eliminate gob fires, dangers from roof falls, and surface subsidence, and also to increase extraction efficiency. From the experience thus gained, it was decided that the most likely system to give maximum benefits, considering the existing geological conditions and coa1-seam characteristics, was hydraulic stowing. This paper examines trends in the working techniques of hydraulic stowing and discusses the favorable results and technical problems that the method produced at Armutcuk. The Armutcuk mine, a subsidiary of Eregli Coal Mines (EKI), produces 1300 tons of run-of-mine coal daily from thick (about 10-90 ft) and highly dipping (about 50-90°) coal seams. The coal is highly volatile with good coking properties that make it highly suitable for industrial use. Most of this coal is used in the iron and steel industry at Eregli on the Black Sea coast. Complete recovery of the coal is of vital importance to industrial development and progress. Initial coal extraction was carried out by means of a sublevel mining system modified to accommodate local geological conditions and coal seam characteristics. This method is similar to that employed in metalliferous mines. Several problems arose from the application of the method, one of them being the prevalence of gob fires which caused periodic shutdowns in some sections. To achieve maximum coal extraction with minimum dilution, to eliminate gob fires and surface subsidence, and to reduce fatal accident rates, a decision was taken to change from this method to a hydraulic stowing system. The pres- ence of large deposits of conglomerate and sandstone on the surface contributed to the decision to employ this new method. Although hydraulic stowing practices have been carried out in America and Germany, Upper Silesia may, so far as coal is concerned, be considered the birthplace of the method. Here the process was adapted to thick coal seams on a large scale to reduce dangers from gob fires and damage from surface subsidence. Today, the mining system is not only widely used in the coal mines of France, the Saar, India and Poland, but has also been successfully applied to the iron ore and potassium salt mines in Russia and Germany. LOCATION AND HISTORY The Kandilli-Armutcuk coal field, the most western outcrop of the North Anatolian coal basin, extends for five miles along the Black Sea coast (Fig. 1). The Armutcuk Mine is situated on the southern boundary of the coal field about 40 miles from Zonguldak and about 12 miles from Eregli. The history of coal mining in the North Anatolian coal basin goes back more than a century. Mining commenced on a small scale in the Kandilli-Armutcuk coal field in the late 1880's, and from this time until nationalization in 1940 the mines were operated by private Turkish and foreign interests. In 1940, Eregli Coal Mines was established to takeover the management of these enterprises. The Armutcuk Mine was developed for the first time in the mid-1940's. Since then it has been in continuous production and it now produces an annual 440,000-ton run-of-mine output from which 320.000 tons of salable coal is obtained. MINE GEOLOGY The carboniferous inlier of the Kandilli-Armutcuk coal field crops out over an area about three miles long and one mile wide. The major part of the inlier consists of the Namurian formation with only a relatively small area of Westfalian-A visible on the surface (Fig. 2). The Namurian formation, which has no exploitable coal seams in the Armutcuk Mine, consists mostly of arkose sandstones and black shales often having

Jan 1, 1969

By Shiou-Chuan Sun

THE fact that coals of different ranks and even of the same rank differ greatly in their amenability to iroth flotation is well known. In recognition of the need for an explanation of this phenomenon, two hypotheses have been suggested. Wilkinsl reported that the floatability of coals increased with an increase of the carbon content or rank. This postulate is handicapped by the fact that bituminous coals that possess moderate carbon contents are actually more floatable than anthracite coals that have high carbon contents, as shown in columns 6 and 9 of Table I. Taggart and his associates' implied that the difference of floatability between bituminous and anthracite coal was caused by the variation of carbon-hydrogen ratio. This is not applicable to the relative floatability of other coals and carbons. For example, column 11 of Table I shows that the carbon-hydrogen ratios of low-floating lignitic coal and non-floating animal charcoal are not only smaller than the moderate-floating anthracite coal, but are also similar to the high-floating bituminous coal. Furthermore, according to this hypothesis, high temperature coke-A (464), Ceylon graphite (1238), and lamp-black (357), all possessing extremely high carbon-hydrogen ratios, should be less floatable than other substances having much lower carbon-hydrogen ratios such as high volatile-B bituminous coal (11.9 to 22), anthracite coal (35.7 to 60.5), lignitic coal (15.6 to 33.6), and charcoal (13 to 26.2). However the former group is actually more floatable than the latter group. In this paper, a surface components hypothesis is Proposed to explain the different floatabilities of coals, carbons, and hydrocarbon minerals. The validity of the hypothesis is experimentally supported by the actual floatability, natural floatability, wettability, and adsorbability for neutral oils of coals, carbons, and hydrocarbon minerals tested. The combustible recovery of the flotation results, as used in this paper. was calculated from Eq. 1: P (100-Ep) 100 RWCP Rc= [1] F (100-E,) C, where R, is the percent combustible recovery; F and P are, respectively, the weight of feed and the weight of concentrate or product; E, and Ep are, respectively, the total percent of ash plus moisture in feed and in concentrate; Ru. is the percent weight recovery: and C, and C, are, respectively, the percent of combustible in feed and in concentrate. Except for ash and moisture content, all chemical components of a coal are assumed combustible. The experimental work included studies on flotation, ultimate and proximate analyses, contact angle tests, extractions of bitumen-A with benzene, adsorptions for liquid hydrocarbons, and wetting tests. Most of the flotation experiments were performed in a laboratory Fagergren machine; others were tested in a small Denver machine. The solid feed for the former was 300 g and for the latter was 30 g. The solid materials used for flotation were crushed to —48 mesh. After the mineral pulp in the flotation cell was agitated for 6 min and the pH was adjusted to 7.5 & 0.2 with sodium hydroxide or hydrochloric acid, a petroleum light oil having a viscosity of 5.73 centipoises at 77 °F was added and conditioned for 2 min. Finally, pine oil was introduced and the froth was collected for exactly 3 min. The weight ratio of petroleum light oil to pine oil was kept constant at 1.5 to 1. Tap water was used for all flotation tests. Contact angles were measured with a captive bubble machine. For each coal sample, three specimens were mounted in transoptic mounts and polished with levigated alumina, first on a sheet glass, then on a cloth-covered metal polishing wheel. The polished specimen was first washed with distilled water and wiped thoroughly on a cleaned linen pad, then transferred into the pyrex cell of the captive bubble machine and conditioned for 6 min., and finally measured for contact angles at three or more points. Except where otherwise stated, the induction time for each measurement was 1 min. The contact angle representing each material was obtained by averaging the measurements of three specimens. The linen pad was first washed with warm distilled water, then boiled 30 min in a 2N sodium hydroxide solution, and finally washed with distilled water until no trace of sodium hydroxide could be detected in the decanted solution. The cleaned linen pad was stored under distilled water. Immediately before using, the pad was rewashed and transferred into a clean pyrex petri dish partly filled with distilled water. The glassware and rubber gloves used were cleaned by soaking in sulphuric acid-potassium dichromate cleaning solution, followed by rinsing with distilled water. The polished specimens were handled only by glass forceps. The ultimate and proximate analyses were made in accordance with the ASTM standard procedures for coal and coke. The extractable bitumen-A was determined by weighing 1 g of —100 mesh sample and placing it in a desiccated and weighed ASTM aluminum-extraction thimble. The thimble was placed in condenser hooks and inserted into an extraction flask containing 100 cu cm of benzene. The flask was heated and the benzene vapor was condensed by water coils. At the end of 24 hr of percolation, the thimble was removed, desiccated, and weighed. Loss in weight of sample was taken as bitumen-A and calculated to dry and ash-free basis.

Jan 1, 1955

By J. D. Price, W. M. Bertholf

Making maximum possible use of available equipment and material, CF&I placed a high-efficiency, high-capacity washery unit in the existing buildings to gain simplified operation, reduced manpower requirements, raised efficiency, and boosted production. THE central washing plant of the Colorado Fuel & Iron Corp. was first operated in 1918 to furnish coal for two 60-oven batteries of Koppers design. Prior to that time the coke for the blast furnaces had been secured from a number of beehive operations, some with their own washeries. The original washery had four Elmore jigs, which did not do as well as had been expected. In 1923 the jigs were taken out and replaced by Deister Plat-O tables, with various changes in auxiliary equipment. In 1936 the throughput of the plant was considerably increased by the installation of vibrating filters for recovery and drying of fine coal, thereby reducing a large recirculating load of fine coal which passed through the screens of the Carpenter driers. In the late 1940's it became evident that something would have to be done to keep the washery abreast of progress at Pueblo and other washeries. Coal requirements were nearly two and one-half times the original requirements of the 120 small ovens, as additional ovens had been built in 1930, 1938, and 1945, whereas no additional coal washing capacity had been added since the 1936 modifications. Much of the equipment, moreover, had seen better days. One answer to the problem was to build a new and completely up-to-the-minute plant, with sep- arate sections for coarse, medium fine, fine, and very fine coal. On the other hand, it might be financially more profitable to place a high-efficiency, high-capacity unit in the existing buildings. The washing problem at Pueblo is not unique, but it is approached differently. There are no established specifications for coke ash and sulphur. The plant is operated on the principle that if it is more economical to wash it out than to slag it out, wash it out; otherwise, send it on to the blast furnaces. With coals at the Pueblo plant costing what they do, the benefits of washing at a low gravity are questionable. Hence the primary problem is to eliminate the high-ash components and retain all the low and medium-ash components. Sulphur is no problem at all, averaging 0.6 pct in the coke. Since a good proportion of the free impurity in the coals is as large as the coal or bone or larger, it appeared that it would be safe to go to jig washing and discontinue crushing the coal before washing, as had been the practice with the tables. This would boost the throughput. It might also decrease manpower requirements and washing costs considerably. It was assumed at the start and has been demonstrated in the first year of operation that a jig will do as well on the fines as the tables had been doing. It would not be necessary to discount any gains due to increased efficiency in washing the coarse material because of incapacity to wash the fines. By washing at —2+0 mesh it would be possible to dewater, by screening, a considerable tonnage satisfactorily, for example, at —2 + % or —2 + l/4, and then crush it in a hammer mill. The —?+20 could be dewatered in centrifugal driers and the remain-

Jan 1, 1955

By H. L. Riley, B. W. Gandrud

AS far as the present writers know, this system of flotation has not been used elsewhere in this country, but in the last couple of years it has been introduced, with minor variations, at one plant in England and one in Wales.' The system has been described and discussed in a number of publications.2-5 The following is quoted from an abstract of the latest of these,5 a paper presented at an International Conference on Industrial Combustion in 1952. On the basis of experience to date with the commercial plants, it is believed that the kerosene-flotation process incorporates all the necessary elements to make it greatly superior to anything else now available for treating of fines in wet processes of coal preparation. Additional study and investigation are still needed, however, to determine if certain phases of the process can be improved to such an extent as to make it generally satisfactory and acceptable to the industry. Further improvements will be needed with respect to the capacities of the flotation cells and the reagent consumption. The situation referred to above explains why an investigation is being made of the possibilities of achieving better cell capacities. Results obtained from this investigation, which is still in progress, are believed significant with regard to both cell capacity in general and the relation of cell design to cell capacity in particular. Commercial equipment now being used in a laboratory-type investigation should have performance characteristics similar to those of the larger machines. Equipment and Procedures: All flotation tests have been made in a standard Denver sub-A 24x24-in. unit cell of 12-cu ft volume. Cell modifications to make it more suitable for the tests were an adjustable front-wall section for varying cell depth and a perforated scraper-drag assembly for removal of the float product. There is also an apron dry-coal feeder, a gravity-feed water supply, reagent feeders, and a centrifugal pump that feeds the mixture of coal, water, and reagents into the flotation cell. A wattmeter connected into the drive-motor circuit records the power requirements of the impeller throughout each run. Dry coal, water, and reagents are all fed through a pan-type intake to the feed pump. A Sturtevant blower was set up to furnish air for supercharging. A centrifugal pump with a garbage-can intake provides for disposal of refuse flow to an outside settling tank. Figs. 1 and 2 show the flotation cell; Fig. 2 also illustrates the blower for supercharging. For purposes of this investigation, the percentage by weight of the feed coal recovered in the float product under a standard set of conditions has been considered as the criterion of cell capacity. The authors realize that such a criterion may be somewhat unorthodox, as the term cell capacity is usually understood to refer to feed input and ordinarily takes into account the ash analyses of the float product and refuse. However, the word capacity is flexible enough so that Webster gives one definition as maximum output, a definition which seems to justify, at least partly, acceptance of the above criterion. It has been the authors' experience in the Birmingham district that the ash-reduction efficiency of the coal-flotation process is generally satisfactory and that the only real problem is to increase the rate of float recovery so that the feed rate to any given bank of cells can be increased without undue loss of coal in the refuse. Originally it was planned to operate the flotation cell to simulate continuous operation during sampling periods. It was assumed that operating for reasonable time with feed coal, water, and reagents turned on would stabilize conditions so that the weight of float coal discharged during a fixed time interval would be an accurate measure of the rate at which the coal was being floated. It developed, however, that this supposition was erroneous. The float coal, caught for fixed time intervals and weighed, gave widely varying results in duplicate runs. Efforts to correct this trouble failed, and it was decided to try to operate on a batch-test basis, whereby all the float coal produced during a run on a known weight of feed coal would be caught in tubs, dewatered, and weighed. This method gives consistent and reproducible results, with total float product weight rarely varying by more than 3 or 4 pct on duplicate runs. The standard test procedure is as follows: A 132-lb sample of dry feed coal is weighed and placed in the feed hopper. The feeder is adjusted for a rate of 800 lb per hr. Feed water and reagents are turned on, and the feed and refuse pumps are started. One minute later the impeller is started. Six minutes are allowed for the cell to fill up with the water-reagent mixture. The feed of dry coal is started at the end of this 6-min period. One minute later the float-coal removal drag is started. The float coal is caught in one tub for the first 6 min after the flow of feed coal starts. Tubs are then changed, and the float coal is caught in a second tub until the feed coal runs out, when the tubs are again interchanged to catch the float coal for the remainder of the run in the first tub. The cell is kept running for 3 min with the water and reagents on after the feed stops to allow residual float coal to be removed. At the end of a test the wet float coal in both tubs is weighed and the total weight recorded. The product in the second tub is used for moisture determination and screen-size analyses. When the

Jan 1, 1956

By Thomas L. Kesler

I all of the 14 million tons of kaolin produced in Georgia through 1949 had been mined from a single deposit 20 ft thick, it would represent a mined-out area of less than 1 sq mile. This measure of depletion is in such strong contrast with the large size of the area shown in Fig. 1, from which most of the output has come, that there has developed an optimistic outlook concerning future supplies of kaolin equal in grade to that already mined. Dozens of mines are widely but not evenly distributed within this area, which consists mostly of parts of Twiggs, Wilkinson, and Washington counties. There are scarcely half a dozen kaolin mines elsewhere in the state that have been appreciably productive, although the areas in which they occur may have good possibilities. Most of the efforts to find new deposits, therefore, will be confined to the area shown in Fig. 1, for it is here that the industry has established its facilities, and prospecting is encouraged by relatively numerous outcrops and opened deposits of kaolin. The Ocmulgee and Oconee rivers drain the area. They and their main tributaries have flat, swampy floodplains covered with alluvium and for the most part thickly wooded. The divides between these streams are intricately dissected by the minor tributaries, which are spring-fed in steep-walled, cirque-like heads. Gullying and under-cutting are active along the streams but are partly controlled by unusually dense vegetation. The mean altitude is about 350 ft above sea level, and locally the relief ranges up to 200 ft. The area is near the inner margin of the Coastal Plain, and is underlain by Cretaceous and Tertiary formations. Its general location is shown in black in Fig. 1. The division between the Coastal Plain and the Piedmont plateau is actually a very sinuous line, the fall line, marking the extent to which erosion has stripped the Coastal Plain sediments from the southeastward-sloping surface of the crystalline rocks of the Piedmont. Cretaceous Host Formation The kaolin deposits, which are described in the succeeding section, occur sporadically in lenslike bodies in a thick series of poorly sorted clayey sands constituting the lowermost of the Upper Cretaceous formations, the Tuscaloosa. The basis for this correlation has been reviewed by Cooke.1 The same formation has been termed the Potomac group by Ladd2 and the Middendorf formation by Smith' in reports on the kaolin deposits. In another report, Veatch4 called the formation the Tuscaloosa but correlated it with the Lower Cretaceous. The Tuscaloosa in this area consists of fine to very coarse sands, commonly cross-bedded, which contain widely different proportions of white, pink, and yellow kaolin and considerable white mica. A few of the coarser beds contain quartz grains up to 3 in. in length, with bodies of pure white kaolin of equal size that appear to have been formed by the weathering of feldspar grains in place. In the kaolin-producing area, gravel and cobbles occur mainly at the base of the formation. Fig. 2 shows the relations of the Tuscaloosa and its kaolin deposits to the adjacent formations. Its base, on the eroded surface of the crystalline rocks, dips southeastward 50 to 60 ft to the mile, according to LaMoreaux.5 The upper surface of the Tuscaloosa is also an unconformity,

Jan 1, 1952

By W. J. Skewes, C. T. Holland

Properly constructed stopes are necessary underground for proper ventilation and safe, healthful working conditions. Mining companies are concerned with providing the best possible stoppings at an economic cost. It is important to consider 1) proper materials, 2) construction that ensures maximum performance from the material used, and 3) keeping costs within reasonable limits for the unit operation—stopping—involved. The authors investigated all aspects of the problem and have attempted to layout a study program to provide information on all phases. Mining companies every year purchase huge amounts of material for the construction of ventilation stoppings. This material by the time it is in place in the stoppings will cost the company many thousands of dollars. In addition and more importantly these stoppings are supposed to confine air travel along certain well defined paths into areas where it is needed to provide safe and healthful atmospheres for workmen. The cost of forcing this air through the mine is large and in general increases as the total quantity increases. Consequently, if the stoppings leak, increased cost results and in addition unsatisfactory ventilation may prevail in the mine. Therefore, it is essentia1 to provide the best possible stoppings at an economic cost. When attempting to provide the best stoppings possible at an economic cost, three aspects of the problem must be considered. These are 1) getting the most suitable material for stopping constructions, 2) constructing the stopping so as to get the best performance from the material used, and 3) keeping the cost of stopping construction to a minimum for a suitable stopping. This investigation is concerned with all phases of the problem and has attempted to lay out a study to provide information on all of them. In order to develop information on the best stopping material available it seemed wise, after con- sidering the job a stopping must do, to investigate the following aspects of material used for stoppings: 1) strength in compression, 2) compressibility or yield under load, 3) porosity, and 4) permeability. To provide information on the best manner of construction the following aspects of air leakage through the stopping were investigated: 1) when the stopping is built of dry stacked materials—i.e., the joints are not mortar-laid (See Fig. 1); 2) when the stopping is built of dry stacked material and then plastered on one side (see Fig. 2); 3) when the

Jan 1, 1961

By L. E. Shiffman

W. J. Parton—Those operators faced with the problem of treating fine coal whether in bituminous or anthracite will find this paper most timely. I would like to take this opportunity of discussing Mr. Schiffman's paper and at the same time express certain views relating to our Tamaqua plant. I would like to ask the author what type of impeller and diffuser is used in the Denver cells? Screen analysis of products from individual cells indicate that coarser material resists flotation and only floats after greater retention time in the last few cells. Also, the need for a scavenger screen to reclaim non-floated coal particles further stresses this point. I have always felt that more efficient means of cleaning coal between 10-mesh and 28-mesh existed than flotation. Reagent and power costs are high for the flotation process. When floating +28-mesh particles, cell capacity is lowered and some of the particles are lost with the refuse. The Tamaqua plant of the Lehigh Navigation Coal Co. floats —28-mesh coal and capacity of recoverable coal is 40 tph for 1800 cu ft of Denver cells; or 0.05 tons per cu ft of cell. At Kimberly 7.75 tph for 600 cu ft of cell gives 0.013 ton per cu ft of cell. At Bessie 14 tph for 800 cu ft of cell gives 0.017 tons per cu ft of cell. It would be appreciated if the author would comment on what he feels is the upper size limit of particle to attain most efficient utilization of the flotation process. The dewatering screw is a very interesting development since it offers a simple way to prepare coal sludge for more complete clewatering by drainage or mechanical dewatering on screen or filters. In other words it could be used to accomplish the same thing as a thickener tank. I would appreciate having the author's comment on how he thinks such a screw dewaterer would work on a froth.* The process as used in floating coal at the Bessie and Kimberly plants may be referred to as more of a bulk oil float in contrast to a froth flotation process. Experiments on increasing capacity of cells is most interesting since we are going through such an experi- mental period at the present time. Recently a double overflow was installed on our No. 30 Denver cells. So far results are not conclusive. In reviewing this paper the following comments are made pertaining to investigation of methods for increasing capacity: Supercharging: Supercharged air in matte flotation or for that matter the use of the normal amount of air drawn in by the impeller would in all probability cause such an aeration in the cell as to destroy the buoyant effect given to the coal particles by the excessive amount of kerosine used. In other words, air creates an agitation zone throughout the cell, creating a boiling and thereby giving a lower recovery in the cell. It would be interesting to know whether the 7 pct increase in recovery was with no air being admitted to the stand pipe. Changing Impeller Speed: The speed of a receded disc impeller for a No. 30 cell as recommended by the Denver Equipment Co. is, I believe, approximately 250 rpm. At this speed and using supercharged air in excess of 8-oz pressure, we have observed a boiling action in the cells. In our flotation we endeavor to obtain some degree of froth flotation using pine oil as a frother. The boiling action as caused by increasing the amount of air added to the cells is detrimental to recovery in froth flotation. It is our belief that to obtain increased recovery from a cell in froth flotation, additional air must be introduced but at the same time this air must be dispersed throughout the pulp in the form of small bubbles and this can only be done by increasing the speed of the impeller. Therefore, if Mr. Schiffman decreased the speed of the No. 30 impellers and at the same time continued to use supercharged air, the boiling action may have been increased because larger bubbles developed. The lower recovery as reported could be due to this factor. Decreasing the impeller speed will definitely decrease the power consumed but may have other disadvantages. First, we believe it will permit "sanding" in the cell and this in our opinion will increase the wear on the impeller and diffuser, especially so, if there is pyrite and/or sand present in the feed. "Sanding" in the cell when air is used, as in froth flotation, will effect the dispersion of this air and cause boiling.

Jan 1, 1951

By L. E. Shiffman

W. J. Parton—Those operators faced with the problem of treating fine coal whether in bituminous or anthracite will find this paper most timely. I would like to take this opportunity of discussing Mr. Schiffman's paper and at the same time express certain views relating to our Tamaqua plant. I would like to ask the author what type of impeller and diffuser is used in the Denver cells? Screen analysis of products from individual cells indicate that coarser material resists flotation and only floats after greater retention time in the last few cells. Also, the need for a scavenger screen to reclaim non-floated coal particles further stresses this point. I have always felt that more efficient means of cleaning coal between 10-mesh and 28-mesh existed than flotation. Reagent and power costs are high for the flotation process. When floating +28-mesh particles, cell capacity is lowered and some of the particles are lost with the refuse. The Tamaqua plant of the Lehigh Navigation Coal Co. floats —28-mesh coal and capacity of recoverable coal is 40 tph for 1800 cu ft of Denver cells; or 0.05 tons per cu ft of cell. At Kimberly 7.75 tph for 600 cu ft of cell gives 0.013 ton per cu ft of cell. At Bessie 14 tph for 800 cu ft of cell gives 0.017 tons per cu ft of cell. It would be appreciated if the author would comment on what he feels is the upper size limit of particle to attain most efficient utilization of the flotation process. The dewatering screw is a very interesting development since it offers a simple way to prepare coal sludge for more complete clewatering by drainage or mechanical dewatering on screen or filters. In other words it could be used to accomplish the same thing as a thickener tank. I would appreciate having the author's comment on how he thinks such a screw dewaterer would work on a froth.* The process as used in floating coal at the Bessie and Kimberly plants may be referred to as more of a bulk oil float in contrast to a froth flotation process. Experiments on increasing capacity of cells is most interesting since we are going through such an experi- mental period at the present time. Recently a double overflow was installed on our No. 30 Denver cells. So far results are not conclusive. In reviewing this paper the following comments are made pertaining to investigation of methods for increasing capacity: Supercharging: Supercharged air in matte flotation or for that matter the use of the normal amount of air drawn in by the impeller would in all probability cause such an aeration in the cell as to destroy the buoyant effect given to the coal particles by the excessive amount of kerosine used. In other words, air creates an agitation zone throughout the cell, creating a boiling and thereby giving a lower recovery in the cell. It would be interesting to know whether the 7 pct increase in recovery was with no air being admitted to the stand pipe. Changing Impeller Speed: The speed of a receded disc impeller for a No. 30 cell as recommended by the Denver Equipment Co. is, I believe, approximately 250 rpm. At this speed and using supercharged air in excess of 8-oz pressure, we have observed a boiling action in the cells. In our flotation we endeavor to obtain some degree of froth flotation using pine oil as a frother. The boiling action as caused by increasing the amount of air added to the cells is detrimental to recovery in froth flotation. It is our belief that to obtain increased recovery from a cell in froth flotation, additional air must be introduced but at the same time this air must be dispersed throughout the pulp in the form of small bubbles and this can only be done by increasing the speed of the impeller. Therefore, if Mr. Schiffman decreased the speed of the No. 30 impellers and at the same time continued to use supercharged air, the boiling action may have been increased because larger bubbles developed. The lower recovery as reported could be due to this factor. Decreasing the impeller speed will definitely decrease the power consumed but may have other disadvantages. First, we believe it will permit "sanding" in the cell and this in our opinion will increase the wear on the impeller and diffuser, especially so, if there is pyrite and/or sand present in the feed. "Sanding" in the cell when air is used, as in froth flotation, will effect the dispersion of this air and cause boiling.

Jan 1, 1951

By G. H. Matheson, J. M. W. Mackenzie

The flotation rate of coal has been studied using a continuous laboratory flotation cell and a multiple exposure photographic technique. The effects of particle size, reagent additions and cell turbulence on the flotation rate of coal have been investigated. The results have been analyzed in terms of current theories of particle bubble encounter and adhesion. The rupture forces which occur during particle bubble encounter have been shown to have an important effect on the flotation rate of coarse coal particles. The flotation of coal is a time dependent process, and in order to obtain a complete understanding of the operation it is necessary to determine the kinetics of the process. In coal flotation, the flotation rate and the associated values of cell productivity in terms of tons per kilowatt-hour and tons per cubic foot of cell volume are of greater importance than in most other mineral flotation operations. The reasons for this are associated with the low unit value of coal and the relatively high cost of coal flotation compared with other coal cleaning processes. An important aspect of coal flotation kinetics is the design of coal flotation cells. The cells used in many coal cleaning operations are often replicas of those used for the treatment of sulfide ores and are not necessarily suitable for the treatment of coal washing pulps, which usually contain a higher percentage of coarse and floatable material than the sulfide pulps for which the cells were designed. This disadvantage of many coal flotation cells has been recognized by Russian engineers who, by the addition of baffles to their subaeration coal flotation cells have increased the cell capacity by 100%. Investigation of the variables which influence flotation kinetics and cell productivity are, therefore, important to the coal industry where flotation is the only accepted method of cleaning the fine (-1/2mm) coal. The kinetics of the flotation process have been investigated in detail by Schumann10 who developed a continuous flotation cell which permitted the study of flotation under steady state conditions. This technique has been used by Gaudin et a1.,6 Brown and Smith and recently by Jowett and Savfi.7 The latter authors have developed a refined mathematical treatment for interpreting the results of continuous laboratory tests. Attempts have also been made to study flotation kinetics using conventional batch testing techniques. The experimental difficulties, however, and the problems of determining the order of the flotation rate process have made this technique of dubious value. A summary of the proposed rate equations for batch testing and the difficulties associated with this technique is given by Brown and Smith. 2 The experimental techniques described above are concerned with the rate of production of froth product. However, several authors, notably Sutherland, 11 Bogdanov1 and Philippoff, 9 have analyzed flotation

Jan 1, 1963

By M. R. Geer, H. F. Yancey, P. S. Jacobsen

Two growing problems confront the preparation engineer—still further restrictions on stream pollution and a greater proportion of fine coal as more and more continuous miners come into use. The de-watering screens, centrifuges, and settling ponds that sufficed a few years ago must often be supplemented by more effective equipment, and in some instances the finest solids are now being recovered by vacuum filtration, once considered too costly in many coal washeries. Unfortunately not all slurries can be readily filtered. Some are virtually unfilter-able, and others do not permit the high cake rates needed to hold costs within reason. Often, however, these difficult slurries can be rendered filterable by flocculation. Starch and lime have long increased settling rate in thickeners and can also aid filtration. More recently there have been other flocculants, primarily synthetic polymers or gums. Reports of their great effectiveness prompted the U. S. Bureau of Mines to test their use in filtration as part of its program on recovering and cleaning fine coal. The object of the present work was to compare some of the newer flocculants with starch and lime and to test the reaction of different types of slurries to flocculation. Three natural slurries, three synthetic coal-clay mixtures, and five flocculants were tested, all with a laboratory filter leaf. Slurries and Flocculants Tested: First of the slurries was the thickener feed from the washed coal section of the Michel colliery of Crow's Nest Pass Co. Ltd., Michel, B.C. As shown by the screen and ash analyses in Table I, practically all this slurry was finer than 28 mesh, and about half the solids were finer than 200. Substantially all material coarser than 200 mesh was clean coal, but the ash content of 20.7 pct in the finest size indicates a moderate amount of impurity. This coal is medium-volatile bituminous in rank. Slurry was also obtained from the Black Diamond washery of Palmer Coking Coal Co., King County, Wash. This underflow of a 1/2-mm vibrating slurry screen is a waste product discharged to a settling pond. As shown by the data in Table I it contained about 20 pct material finer than 200 mesh, which analyzed 77.5 pct ash, indicating a high proportion of clay. The coarser sizes were also high in ash content because of the presence of bone and shale. Black Diamond coal is on the dividing line between bituminous and subbituminous rank. The third slurry, from the washery of Roslyn-Cascade Coal Co. in Kittitas County, Wash., was the underflow of a battery of 8-in. cyclones that is wasted with the coarser refuse from the plant. In both size and ash content this slurry was similar to the Black Diamond, but since it was a cyclone-underflow product, the fraction finer than 200 mesh undoubtedly contained a much smaller proportion of finely divided clay than was present in the Black Diamond slurry. The Roslyn-Cascade coal is high-volatile A bituminous in rank. Initially the slurry samples were stored in the laboratory at a concentration of 40 pct solids. A progressive change in filtration and flocculation characteristics during storage was noted, however, presumably because of progressive disintegration of the clay. After careful testing demonstrated that the slurry solids could be dried and then repulped without changing the original filtration characteristics, this procedure was adopted. The flocculants used were Separan 2610, Kylo 27, Jaguar MD-A, Idaho potato starch, and lime. The lime was used as a slurry, the Separan 2610 as a 0.10 pct aqueous solution, and the other flocculants as 0.50 pct aqueous solutions. The starch was causti-cized with sodium hydroxide before use. As some of the flocculants are reported to deteriorate on storage, new batches of flocculants were prepared every two days. Test Procedure: Fig. 1 diagrams the filter leaf apparatus employed. This consists of a test leaf having an effective area of 0.10 sq ft., a 60°, 8 1/2-in. diam cone to hold the slurry sample, a variable-speed electric stirrer, vacuum gage, bleed line for controlling vacuum, filtrate flask, moisture trap, and vacuum pump. Two filter cloths were selected—Saran SA 603, a monofilament, coarse-weave cloth, and polyethylene PO-801 HF, a monofilament, fine-weave cloth. The Saran was selected to provide maximum cake rate without regard to clarity of filtrate and the polyethylene to provide the best quality of filtrate without regard to cake rate. The test charge for the filtering vessel consisted of 800 g of solids (previously soaked for 24 hr) and

Jan 1, 1960

By Frank Doxey

AT Dominion Steel & Coal Corp. it has long been recognized that continued mechanization of mine operations is necessary in the Pictou, Cumberland, and Sidney coal fields of Nova Scotia. The varied physical conditions in these fields call for special consideration of individual cases before planning is finalized. Because standard equipment cannot be procured which would operate successfully, many experiments have been necessary over the years to keep pace with the progress made in other countries. There are two mines, producing 2000 tpd, located in the Pictou coal field. The field is badly distorted and crossed by many faults. Seams are highly inclined and irregular and vary in thickness from 5 to 40 ft. Entries are difficult to maintain because of squeezing of the coal. ribs and movement of the roof and pavement. Output from the three operating mines in the Cumberland field is 3000 tpd. The field is highly inclined, inclination varying from 12° to 32°. Overlying beds consist of shales and massive sandstone lenses of extreme toughness and are responsible for bumps when the stresses are relieved by extraction. At greatest depth these are among the deepest coal workings in the world. Depth of cover ranges from 2300 to 4000 ft. This prohibits room-and-pillar working and necessitates longwall operation. Working of contiguous seams concurrently to maintain output increases the already difficult conditions. The Sydney field, with a frontage of about 30 miles, is the most important of the Nova Scotia coal fields. With the exception of one small area it is now wholly submarine. Output is approximately 21,000 tpd. Seams are 21/2 to 8 ft thick, and cover in the areas varies from 600 to 2300 ft, with an average of 1500 ft from sea bottom to the top of the seam. The seams dip in a seaward direction, pitches ranging from 6" to 4.0". The shoreline is the last place of entry to the seams and distance from the bank to the working faces is generally over 3½ miles, in some cases as much as 6½ miles. Ventilation is a problem and requires the construction of large permanent airways. Getting and Loading Coal: In 1925, in view of heavy pressures exerted by thickness of cover overlying the seams, roadways and pillars of the room-and-pillar system being worked began to break up and coal was lost. It was decided that a change in the method of extraction in areas with heavy cover was necessary, and experiments were made with many short walls and longwalls varying between 90 and 250 ft. Trial and error proved that the best operating length was between 400 and 500 ft, delivering all coal to the dip, with roadways to the face following a level course. The change-over was gradual, and the technique of roof control developed with the system, so that falls to the face are now very infrequent. An advantage of longwall mining is that it yields 95 pct extraction, especially important in coal seams of high quality or in seams where faults or disturbances restrict the workable areas. This percentage of extraction is based on the fact that the longwall advancing system takes development faces where pitch permits instead of driving headings and leaving roadway pillars. This system yields high tonnage during development and limits loss in extraction to duff left during operations. Accompanying disadvantages, on the other hand, are the heavy construction cost of main roadways and the necessity of driving all new flank face roadways through the gob. If the main roadways are driven through the solid, and large enough pillars are left on each side for protection against flank face weights, then the width of solid coal is approximately 1700 ft. This represents 10.6 pct of the coal available, or 89.4 pct extraction of the whole. These pillars, however, are of such size that they provide a useful pillar drawing area as a final operation of the mine. It may be that although the seam cannot be generally mined by the room-and-pillar method, it can be adapted to

Jan 1, 1955

By R. R. Estill

THE rank of the Ruhr coal ranges from a high volatile bituminous coal to an anthracite, depending to some extent on the original depth of the seam. The average Ruhr coal corresponds to a soft bituminous American coal of a coking quality. The average thicknesses of individual coal seams being mined are also comparable (59 in. against 65 in. in the United States). However, consideration of seam conditions and mining conditions other than those just mentioned emphasizes differences rather than similarities with United States soft coal. In general, the Ruhr seams now being mined are much more folded and inclined than American seams. Dips of 20' and 30" are common in seams now being worked, and 30 pct of the coal reserves in the district are in seams dipping more than 35". Only on the tops and bottoms of folds do we find rather flat coal seams. In addition to the folding there is extensive displacement by cross faulting plus a certain amount of strike faulting of an overthrust nature, which results locally in doubling or omission of seams. Because of the long history of mining in the Ruhr, nearly all coal lying near the surface has long since been mined out, and we find that the average depth of mining is at present about 2300 ft below the surface. Deep mining, folding, and faulting result in seam conditions requiring a great deal more roof support than one finds in American soft coal mines. In fact only in the anthracite district and the Rocky Mountain and Pacific coal fields do we find somewhat similar conditions. It is easy to say, therefore, that the problem of mechanization of coal cutting and loading in the German mines is quite different from that which we have so effectively met in America with our mobile cutters and loaders, duck bill loaders, and a room and pillar system of mining our drift and slope mines. Partly because of more limited coal reserves, the traditional German mining system is largely the longwall method, which gives an almost complete coal recovery. Backfilling must be extensively practiced to protect the longwall faces, the over and underlying seams and workings, and especially the surface industrialized areas and barge canals. The German engineers have accordingly concentrated their efforts on the design of cutters, loaders, and conveyors suitable to longwall methods rather than room and pillar methods. Undercutters with cutter bars like American models have been in use in the Ruhr since well before World War 11. In 1941 they accounted for 8.5 pct of the production. This percentage, of course, includes coal which was undercut but nevertheless had to be broken down with air hammers or with explosives. The most common of these cutters is the Eickhoff Standard cutter (see fig. 1). This machine does about 95 pct of the undercutting in the Ruhr today, and is available with either compressed air or electrical power and in at least four different sizes. A variation of the cutter is this one with two cutter bars (fig. 2). At the end of 1947 about 200 of these machines and similar cutters were accounting for 13.2 pct of the total production, a production which was, however, only 60 pct of the 1941 production rate, so that the actual cutter tonnage was only up to a small amount over 1941. In 1941 about 3 pct of the production was accounted for by shearing machines making their cut perpendicular to the longwall face. They were similar to those used in the States. These machines are today considered obsolete and now account for only 0.7 pct of the total production. They are located at only a few mines and at present do not seem to have much of a future in the Ruhr. For the future, the Ruhr miner is looking forward to rather extensive mechanization of face work, with two major types of equipment being developed almost simultaneously. On one hand there is the development of cutter loaders for use in relatively hard coal. They represent the further extension of ideas developed after relatively long experience with the Eickhoff cutter. On the other hand there has been since 1942 an intense interest in the Ruhr in the development of face-stripping methods, particularly by the Kohlenhobel (coal plow) and its modification. At the end of 1947 these cutter loaders, Kohlen-hobels and scrapers together were actually accounting for only about 1.4 pct of total production while air hammers still broke 77.1 pct and as much as 1.2 pct was actually broken by hand picks. However,

Jan 1, 1951

By R. R. Estill

THE rank of the Ruhr coal ranges from a high volatile bituminous coal to an anthracite, depending to some extent on the original depth of the seam. The average Ruhr coal corresponds to a soft bituminous American coal of a coking quality. The average thicknesses of individual coal seams being mined are also comparable (59 in. against 65 in. in the United States). However, consideration of seam conditions and mining conditions other than those just mentioned emphasizes differences rather than similarities with United States soft coal. In general, the Ruhr seams now being mined are much more folded and inclined than American seams. Dips of 20' and 30" are common in seams now being worked, and 30 pct of the coal reserves in the district are in seams dipping more than 35". Only on the tops and bottoms of folds do we find rather flat coal seams. In addition to the folding there is extensive displacement by cross faulting plus a certain amount of strike faulting of an overthrust nature, which results locally in doubling or omission of seams. Because of the long history of mining in the Ruhr, nearly all coal lying near the surface has long since been mined out, and we find that the average depth of mining is at present about 2300 ft below the surface. Deep mining, folding, and faulting result in seam conditions requiring a great deal more roof support than one finds in American soft coal mines. In fact only in the anthracite district and the Rocky Mountain and Pacific coal fields do we find somewhat similar conditions. It is easy to say, therefore, that the problem of mechanization of coal cutting and loading in the German mines is quite different from that which we have so effectively met in America with our mobile cutters and loaders, duck bill loaders, and a room and pillar system of mining our drift and slope mines. Partly because of more limited coal reserves, the traditional German mining system is largely the longwall method, which gives an almost complete coal recovery. Backfilling must be extensively practiced to protect the longwall faces, the over and underlying seams and workings, and especially the surface industrialized areas and barge canals. The German engineers have accordingly concentrated their efforts on the design of cutters, loaders, and conveyors suitable to longwall methods rather than room and pillar methods. Undercutters with cutter bars like American models have been in use in the Ruhr since well before World War 11. In 1941 they accounted for 8.5 pct of the production. This percentage, of course, includes coal which was undercut but nevertheless had to be broken down with air hammers or with explosives. The most common of these cutters is the Eickhoff Standard cutter (see fig. 1). This machine does about 95 pct of the undercutting in the Ruhr today, and is available with either compressed air or electrical power and in at least four different sizes. A variation of the cutter is this one with two cutter bars (fig. 2). At the end of 1947 about 200 of these machines and similar cutters were accounting for 13.2 pct of the total production, a production which was, however, only 60 pct of the 1941 production rate, so that the actual cutter tonnage was only up to a small amount over 1941. In 1941 about 3 pct of the production was accounted for by shearing machines making their cut perpendicular to the longwall face. They were similar to those used in the States. These machines are today considered obsolete and now account for only 0.7 pct of the total production. They are located at only a few mines and at present do not seem to have much of a future in the Ruhr. For the future, the Ruhr miner is looking forward to rather extensive mechanization of face work, with two major types of equipment being developed almost simultaneously. On one hand there is the development of cutter loaders for use in relatively hard coal. They represent the further extension of ideas developed after relatively long experience with the Eickhoff cutter. On the other hand there has been since 1942 an intense interest in the Ruhr in the development of face-stripping methods, particularly by the Kohlenhobel (coal plow) and its modification. At the end of 1947 these cutter loaders, Kohlen-hobels and scrapers together were actually accounting for only about 1.4 pct of total production while air hammers still broke 77.1 pct and as much as 1.2 pct was actually broken by hand picks. However,

Jan 1, 1951

By A. Roberts

This paper reviews progress in equipment design and in the techniques applied to mine lighting in Great Britain. The incidence of the disease miner's nystagmus, which was a major source of concern only a few years ago, has now diminished, largely owing to improvements in standards of lighting in mines. The reorganisation of the coal mining industry, involving increasing mech-anisation, together with new safety legislation, has brought about the widespread use of mains electric lighting underground. An account is given of some of the results of research to determine the lighting conditions needed for compliance with the compulsory legal requirement of "sufficient and suitable" general lighting at specific localities in British mines. Some details are given of the techniques applied and the results of recent research in the problem of glare from miners' cap lamps. Of the various factors in the working environment of the miner, it is true to say that lighting is seldom given attention comparable with that devoted to other factors, such as dust, heat and humidity, or atmospheric gas content. No doubt one reason for this lies in the ability of the human eye to adapt itself to low brightness levels. Be that as it may, the fact is that at the turn of the 20th century the miner's working light improved little throughout the centuries during which mining technology has developed. Indeed, in coal mines which encounter a methane explosion hazard, flame safety lamps, with a luminous intensity less than one candle power, were virtually the sole means of lighting during almost the whole of the 19th century. The 20th century has witnessed a substantial improvement in lighting techniques underground in coal and metalliferous mining. In Great Britain the general pattern has been one of standards periodically raised by the implementation of Statutory Regulations imposed under successive hlines Acts, but it is fair to add that the more progressive mining companies usually anticipated the Statutory Regulations, which when they have been put into effect have served to turn into general practice what was already being done at the better mines. In earlier years, official requirements were limited to the specification of approved lamp designs, with stipulated minimum values of luminous intensity and light output, but more recently the necessity of satisfactory maintenance of standards during service has been recognized, and the current regulations not only require that a maintenance organisation should exist, but also specify a maximum permissible depreciation below initial approval standards.'

Jan 1, 1961

By H. J. Modi, D. W. Fuerstenau

pH effectively regulates the flotation of corundum through its control of the surface charge. Since collector ions function as counter ions in the double layer, the collector must be anionic when corundum is positively charged and cationic when it is negatively charged. Even a long-chained xanthate will float positively charged corundum. Association of the hydrocarbon chains of the adsorbed collector ions gives rise to increased adsorption and increased rate of flotation. Because of this phenomenon, neutral organic molecules aid flotation although the unionized collector cannot promote flotation independently. Na2SO4 was found to depress the flotation of corundum with animic collectors through ionic competition whereas the same salt activates corundum for flotation with cationic collectors. A close connection between ion association, electrokinetic behavior, and floatability is suggested. As part of the increasing interest in the physical chemistry of flotation, strong experimental evidence has been obtained which indicates that organic ions are adsorbed as counter ions by non-metallic oxide minerals mainly because of the electrical forces which exist at interfaces(1-3). Thus flotation of such minerals without activation should then be appreciable only when the mineral and the collector are oppositely charged. Since only limited work has been reported on the effect of surface charge upon flotation behavior(3) this general hypothesis on the flotation of insoluble oxide minerals has not been tested thoroughly. Investigating the flotation of a mineral for which extensive electrokinetic data exist yields a way to test this hypothesis more extensively. The electrokinetic behavior of corundum in solutions containing both organic and inorganic electrolytes has been studied quite thoroughly' Electro-kinetic measurements have shown that in the case of corundum, simply by controlling pH the solid surface can be charged either positively or negatively. For this reason corundum was chosen for study. The object of this paper is to present a study of the flotation of corundum and to explain the results in the light of our knowledge of adsorption mechanisms gained from streaming potential investigations. This paper is concerned mainly with the effect of surface charge on flotation with anionic and cationic collectors, but in addition such flotation phenomena as collector hydrolysis, Van der Waals attraction between hydrocarbon chains, ionic competition, and activation are discussed. While the present paper is confined to a reporting of the data obtained and some of the more immediate conclusions, it will be seen that the results, in combination with those on other systems, do promise a

Jan 1, 1961