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Joseph William Richards, was born in Oldbury, Worcestershire, England, on July 28, 1864, of English-Scotch parents, Joseph and Bridget (Harvey) Richards. In 1871, he came to this country with his parents. He attended the public schools of Philadelphia, Pa., where his father was a manufacturing metallurgist, and in 1882, graduated from the Central High School with the degree of A. B. The same year he entered Lehigh University, from which he received the degree of A. C. (Analytical Chemist) in 1886; M. S. in 1891, and Ph. D. in 1893. His connection with the university was continuous from 1887 until his death, and he therefore adds another to the list of names of men Prominent in the A. I. 11. E. who have helped to develop Lehigh University to the important position which it now occupies in mining and metallurgical education in respect of world-wide fame. It is fitting to recall in this connection that, Eckley B. Coxe, a fonnder of the A. I. M. E., four times its vice-president,, and president in 1878-79, was literally a pillar of the university for several years prior to his death; Thomas M. Drown, founder, vice-presi(lent, secretary for 11 years, honorary member, and president in 1897, was president of Lehigh University from 1895 to 1904; and Henry S. Drinker, founder, manager, vice-president, and honorary member, was president of the university from 1905 to 1920. Doctor Richards was no less active in both institutions. With the exception of the years 1886 to 1887, when he was superintendent of the Delaware Metal Refinery of Philadelphia, his professional life was devoted to the upbuilding of the metallurgical department of the university, in part through his continuous services as a member of the teaching staff, and in part through the international renown won for the department by his achievevements as author, investigator, expert and rerognizcd authority in metallurgy, especially the metallurgy of aluminum, metallurgical-chemical calculations, and electrometallurgy. He was assistant instructor in metallurgy and blowpiping in 1887; instructor, 1888-90; instructor in metallurgy, mineralogy and blowpiping, 1890-97; assistant professor, 1897--1903; professor of metallurgy, 1903-1921. He died at his home in Bethlehem, Pa., on Oct. 12, 1921, of heart failure. To the A. I. M. E., for a continuous period of a dozen years, his services were equally loyal and effective. He was vice-president, 1910-11, and 1916-17; councilor, 1912; dirctor, 1913-15; chairman of the Iron and Steel Clommittee, 191-1-21; he was one of the framers of the Insti-

Jan 1, 1922

As described in the section on the Palabora twin crushers, the Palabora ore consists of two types: a carbonatite containing a low titanium magnetite, and a foskorite containing high titanium magnetite. As the Palabora pit was widened and deepened, more and more of the foskorite had to be mined. The 25,000 mt expansion of the Palabora plant which went into operation in early 1977 therefore takes its feed from the foskorite section of the orebody. For the expansion, a third 54 in. by 74 in. gyratory crusher was installed to crush feed for two 32 ft by 15 ft (9.75 by 4.6 m), 9000 hp (6710 kw) autogenous single -stage grinding mills. These units are located near the two existing primary crushers, and the ground ore is pumped 5277 ft (1640 m) to the expanded flotation plant by five stages of centrifugal pumps. The new gyratory is installed in a unique Figure-8 shaped enclosure with the crusher in the smaller 27 ft 11 in. (8. 5 m) diameter cylinder and the maintenance area in the 36 ft 1 in. (11 m) diameter cylinder. Cylinder walls are only 23.6 in. (600 mm) thick because their circular design permits all the concrete to be in compression. A rectangular dump pocket which sits on the smaller cylinder contains two manganese steel faced splitters to protect the crusher spider and countersunk water spray headers under the dumping positions for dust suppression. The dump pocket holds 330 st or 300 mt and is also lined with manganese steel. The crusher is mounted on a reinforced concrete diaphragm over the 363 st (400 mt) crushed ore pocket. In the back of the pocket is plastic pipe. Radioactive level control sources are lowered into this pipe from the crusher support floor. The pipe is protected by concrete covered with steel castings. Level control detectors are mounted on the front wall of the pocket. Liner plates on the crushed ore pocket walls are removed and in- stalled by the use of a circular monorail and trolley mounted on the pocket ceiling. The door in the front of the crushed ore pocket is hinged at the top rather than at the sides and is raised by an electric winch. For removal of the eccentric, rails are installed in the pocket and a unique hydraulically operated cart is moved under the eccentric. A platform is raised to receive the eccentric and lowered so that it may be removed. Crushed ore is removed from the pocket by a hydraulically powered, 8 ft

Jan 1, 1978

Written contribution by J.F. Curtist The paper is an interesting exercise in the application of mathematics to a theoretical situation presumed to oc¬cur underground. However, some of the assumptions on which the calculations are based are unlikely to occur in any actual mining situation. Assumptions made in the paper Objections to these assumptions are noted seriatim. (1) `it is assumed that any loading-rate effect will be small and can be ignored' (p. 3). A loading rate of the order of 50 kN-12 per cent of the maximum load-is customary to prevent the props being dislodged by the blast. (2) `it is unlikely that all the props will be installed with their full stroke available, but for the purpose of this analysis, it is assumed that this is the case' (p. 3). A more reasonable assumption would be that, with the most careful selection of prop length and the most rigid control of stope width, the available stroke length would on average be reduced by the order of 70 mm on installation-20 per cent of full stroke. (3) `each support unit supports a block of 3,3 m2 bound¬ed by vertical fractures with zero cohesion that are parallel and at right-angles to the stope face respec¬tively' (p. 5). Joughin and Jager have shown this to be an unwar¬ranted assumption in respect of the positioning, the direction, and the dip of the fractures present in a stope in rockburst conditions. If, as assumed, there was `zero cohesion' between such blocks, the removal of a hydraulic prop in row 3 supporting such a block would ensure that it fell out before a profile prop could be installed in the same position. (4) `the height of each of these blocks is 3,0 m, the volume of each block is 9,9M3, the mass of each block is 27 225 kg' (p. 5). The concept of individual props being affected by dis¬crete blocks of known and uniform dimensions of relatively low ground velocities (less than or equal to 2,16 m/s) acting in line with the axis of the prop is one that lends itself to neat mathematical calcula¬tions. It does not, however, bear much similarity to the actual conditions that are exposed underground in excavations where rockbursts occur. Ortlepp and co-workers concluded that `the impulse of the shockwave can cause the displacement of the

Jan 1, 1991

By N. M. Raju, S. K. Dutta, M. A. Khan, B. Singh

The paper seeks to present the application of ground control research at Mosaboni Copper Mines to increase the size of stoping areas. Room and Pillar stopes were experimented for increased production dispensing with the conventional breast stoping. The experimental stopes, with spans varying from 10 to 15 m, were extensively instrumented to study the behaviour of wall rocks (chlorite schist) and to determine the optimum span in the workings, dipping at 25° to 45° and 3 to 5 m in height. The observations confirmed that 15 m wide span could be safely worked with 3 m wide rib pillars at 600 to 800 m depth by using rock bolts. Ground conditions improved with the use of full-column cement grouted bolts dispensing with timber supports and the productivity was improved.

Jan 1, 1981

By M. I. Brittan, E. J. J. van Vuuren

Computer analysis, modelling and optimization of gold recovery plants of the Anglo American Group by M. I. BRITTAN M.Sc. (Eng.), M.S., Ph. D., (Member). and E. J. J. van VUUREN M.Sc., (Visitor). Multiple correlation and regression techniques are used to analyse gold plant operating records. Key variables m the process which Influence the gold recovery are Identified and assembled into a mathematical model. This constitutes a valuable tool for the optimisation of variables such as the rates of reagent addition and for assessing the economics of extending the plant. Substantial short term economic benefits have been realized. The techniques provide a rapid diagnostic service without interfering with normal plant operation. Quantitative results are presented and their implications examined. The value of the method is discussed.

By Peng Cao, Jimin Su, Yang Li

"Abutment pressure distribution is different when a longwall panel is passing through the abandoned gateroads in a damaged coal seam. According to the geological conditon of panel E13103 in Cuijiazhai Coal Mine in China theoretical analysis and finite element numerical simulation were used to determine the front pressure distribution characteristics when the longwall face is 70 m, 50 m, 30 m, 20 m, 10 m, and 5 m from the abandoned gateroads. The research results show that the influence range of abutment pressure is 40 m to 45 m outby the face, and the peak value of front abutment pressure is related to the distance between the face and abandoned gateroads. When the distance between the longwall face and abandoned gateroads is reduced from 50 m to 10 m, the front abutment pressure peak value kept increasing.When the distance is 10 m, it has reached the maximum. The peak value is located in 5 to 6 m outby the faceline. When the distance between the longwall face and abandoned gateroads is reduced from10 m to 5 m, the front abutment pressure sharply decreases, the intact coal yields and is even in plastic state. The peak value transfers to the other side of the abandoned gateroads. The research results provide a theoretical basis for determining the advance support distance of two gateroads in the panel and the reinforcement for face stability when the longwall face is passing through the abandoned gateroads.INTRODUCTIONFront abutment pressure is the result of rearrangement of in situ rock stress caused by the mining (removal) of the coal seam. Research on the distribution character of face abutment pressure have significant importance in confirming the width of the coal pillar and the distance of advance support in roadways (Liu et al., 2011; Liu, Jiang, and Zhu, 2015). Currently, there is a lot of research on abutment pressure characteristics of traditional longwall coal faces, thick coal seam longwall faces, isolated faces, and extra wide face abutment pressure distribution (Xie et al., 2006). However, theoretical analysis and research on the distribution law of the fully mechanized coal face to abandoned gateroads are relatively less. Abandoned roadways are often from less mechanized or room and pillar mines. Due to the existence of these roadways pre-longwall mining, the in situ stress has already been redistributed around the roadway (Liu et al., 2007). Thus, when the longwall approaches and mines through the roadway, the abutment pressure distribution behaviour has different characteristics than those of traditional solid undisturbed coal. Especially when the face approaches the abandoned roadway, the understanding of face abutment pressure distribution has significant importance to the timing of support installation (Ren and Ning, 2014). This prevents the occurrence of a roof and rib failure ahead of and on the longwall and highly efficient advance of the longwall safely and productivity. This paper analyses the E13103 face of the Cuijiazhai Mine of Wenzhou Mining Co., Ltd of Kailuan Group, as the engineering case study, and uses FLAC3D numerical simulation software to conduct numerical simulated analysis on the abutment pressure distribution mechanism of longwall coal mining face."

Jan 1, 2018

By Alice S. Allen, James Paone

SURFACE SUBSIDENCE IN MINING The consequences of subsidence are becoming progressively more serious as the consumption of an ever- increasing quantity of minerals conflicts with the needs of an expanding population for surface land area. The inevitable result will be an increase in the amount of land area on which undermining is restricted because surface subsidence must be prevented. Designation of land areas that may be undermined, provided that subsidence is controlled and kept within specified limits, also will increase. The decisions as to which land is to be used for urban, agricultural, or other purposes clearly involve considerations that go far beyond the scope of mining technology. Under the authority assigned to the US Bureau of Mines (USBM) by the Organic Act (May 16, 1910) and its succeeding amendments and pursuant regulations (30 U.S.C. 1-1 1 ), the Bureau conducts scientific and technologic investigations concerning mining and its related problems. Subsidence-control demonstrations that have been conducted under this authority include projects in Wyoming, West Virginia, Illinois, and the Pennsylvania anthracite region. The problenls for technology arc to devise methods for extracting minerals with controlled subsidence in varying geologic settings and for calculating or predicting the amount, extent, and characteristics of the subsidence that accompanies the extraction of our principal minerals mined by high-tonnage methods so that logical choices can be made from the possible alternative approaches. Where underground minerals and fuels are mined and removed, the voids that are created underground generate strong imbalanced stresses in the surrounding and overlying rock strata. The resulting readjustments in the rock masses may cause subsidence of the ground surface. Subsidence implies vertical collapse, because the most conspicuous component of movement is downward. However, the downward component is accompanied by differential horizontal strains that may be more damaging to man-made surface structure.; than the more apparent vertical displacements. SUBSIDENCE CONTROL The most widespread method of alleviating potential subsidence problems in undermined areas has been to backfill mine voids with mine refuse or some other in- expensive material that provides lateral support to the remaining mine pillars and vertical support to the mine roof and overburden. Most USBM backfilling work has been conducted jointly with the Pennsylvania Department of Environmental Resources in the anthracite region of northeastern Pennsylvania. These joint projects were conducted under the authority of the Anthracite Mine Drainage Control Act of July 15, 1955 (Public Law 84-162, as amended), and the Appalachian Regional Development Act of 1965 (Public Law 89-4. as amended). In addition to the Appalachian and mine-drainage projects, the Bureau has either conducted or participated in demonstration projects to develop the "pumped- slurry" method of backfilling mine voids. Three of these projects were conducted in Rock Springs, WY. The first was a field test of the pumped-slurry technique, conducted in 1970 under the combined participation of the City of Rock Springs, the US Department of Housing and Urban Development, the Dowell Div. of the Dow Chemical Co., and USBM (Candeub, Fleissig, 1971). The objective was to demonstrate that a large quantity of sand could be hydraulically injected under pressure through a single borehole, and that filling of the mine voids would be essentially complete. In the earlier "blind-flushing" methods, involving sluicing material through boreholes by gravity, quantities per injection hole ranged from 15 to 765 m3 (20 to 1000 cu yd), and the mine voids were only partially filled. In the first test of the pumped-slurry method, approximately 14 900 m (19,500 cu yd) of sand were injected success- fully through a single borehole. Subsequent information obtained through 43 monitoring boreholes indicated that mine voids below 11 330 m2 (2.8 acres) had been filled. As a result of the successful initial test of the pumped-slurry process, USBM conducted additional backfilling demonstration projects using the new technique. The first full-scale demonstration was carried out in the Green Ridge section of Scranton, PA, between 1972 and 1974. This project proved the feasibility of using the new hydraulic-injection technique to back- fill dry mine voids, as well as flooded voids, and it demonstrated that crushed anthracite refuse could be used as easily as sand in the process. Stabilization was pro- vided for about 202 300 m2 (50 acres) of Scranton, having a population of approximately 1000. Additional demonstration projects at Rock Springs, WY, between 1973 and 1975, resulted in the stabilization of about 364 200 m2 (90 acres). The work was necessary to preserve the physical and economic well- being of the city, because there have been numerous occurrences of subsidence during recent years. A demonstration project completed in 1976 in Rock Springs brought the total stabilized area in that city to 647 500 m2 (160 acres). At a cost of about $3,000,000, a population estimated at 7000 persons and property values exceeding $18,000,000 were protected. Presently. USBM is participating in subsidence- control projects involving backfilling in Pennsylvania, Illinois, and West Virginia. Eleven demonstration projects were in progress by USBM in 1977-eight in the Pennsylvania anthracite region and three in bituminous areas of West Virginia (one) and Illinois (two). The estimated total property value is over $100,500,000, and the total cost of the projects is expected to be nearly $20,000,000. The demonstration projects are designed to adapt the pumped-slurry technique to a variety of subsurface conditions, to increase efficiency, and to reduce costs.

Jan 1, 1982

By P. J. A. Beukes, C. C. La Grange

INTRODUCTION Some years ago the authors proposed a modified procedure1 for carrying out micum index determinations2. The modifications consisted of using a drum of 50 cm internal length, i.e. half the length of the standard micum drum (a modification also adopted by the B.S.J.3), and the use of a + 25 mm coke instead of + 60 mm coke for the test. The proposed modified test and formulae for converting results obtained from it to standard micum test results and vice versa were based on the results obtained during the investigation of about 200 test cokes made in ovens of the South African Steel Industrial Corporation Limited (Iscor), the Fuel Research Institute co-operating with the experiments. Extensive testing of cokes has since been done at the Fuel Research Institute using both the standard and the modified procedures. The numerous test results obtained during this trial period fully support and strengthen the confidence originally expressed by the authors in the modified testing procedure. The scope of the modified micum index determination has in the meantime been extended by applying the testing procedure also to test cokes made in the Institute's experimental coke ovens, and also in this case the modified procedure has proved its usefulness. A difficulty is that when comparing the characteristics of cokes made in the Institute's two experimental coke ovens with each other and with those of cokes made from similar blends in commercial coke ovens, for example those of Iscor, the index values obtained invariably do not check exactly due mainly to inherent differences between the different types of oven4. For instance, resistance to abrasion of a coke made in Iscor's ovens is always appreciably higher than that of a coke made from the same blend (and having a similar moisture content) in the experimental ovens. This makes direct comparisons and the prediction of quality to be expected from commercial products, based on experimental coke oven results, difficult. It may also be added that workers in Great Britain some years ago reported a similar experience5. In the authors' opinion the main reason for the discrepancy is the relatively low height of the experimental ovens resulting in a much lower static pressure on the charge during coking than that obtaining in a commercial oven. The difficulty was pointed out and briefly discussed in a publication4 which appeared at a time when the micum test for coke evaluation had only just been introduced at the Institute, so that only B.S. shatter and abrasion test results were available for discussion in the publication. Needless to say, the introduction of the micum test by no means solved nor even alleviated the problem. On numerous occasions during the past few years the Institute has had to conduct contract investigations in its experimental coke ovens on behalf of companies which invariably desired an indication of the quality of coke likely to be obtained from certain coals or blends when coked commercially, more particularly in Iscor's type of coke ovens. It was, therefore, essential to accumulate statistical data which would enabe the prediction to be readily made. Fortunately the compilation of comparative coke-quality index values for the three types of oven under consideration has become possible as a result of coking investigations conducted in co-operation with Iscor over a number of years. During this period Iscor, on numerous occasions, kindly made available to the Institute portions of coals and blends coked in the Corporation's coke ovens, thus enabling parallel coking tests to be carried out in the Institute's experimental ovens. In this way enough statistical data became available to establish relationships between coking results obtained in Iscor's ovens and results obtained on similar coal charges in the Institute's ovens. The most practical and acceptable method of overcoming the problems of finding a suitable and simple basis enabling direct comparison between the cokes of the different origins mentioned to be made, is the assignment of descriptive arbitrary quality ratings, as indicated below, to the cokes obtained, it being argued that the quality rating assigned to coke made from a given coal or blend in the different coke ovens under standardized conditions should be the same, irrespective of differences in the index values obtained when subjecting the cokes to the usual coke evaluation tests. The system, which should enable the sponsors of coking investigations in the Institute's experimental ovens to make their own interpretations of the results obtained, is explained below. ARBITRARY QUALITY CLASSIFICATION OF COKES The Ml0m index1 (percentage material smaller than 10 mm after the test) of a blast furnace coke is a measure of its tendency towards breeze formation during handling -a low value indicating a low tendency. This is generally of greater importance under ruling South African conditions, than the M'40 index1 (percentage material larger than 40 mm after the test), which can be regarded as a measure of resistance to shatter. In fact, if the M10m value of a coke as made in Iscor's ovens is higher than about 11, the tendency of the coke to form breeze during handling and in the blast furnace is relatively so high that hardly any notice need be taken of its M'40 value, even if this is comparatively high. For this reason, therefore, much more weight should normally be assigned to Ml0m indices in the quality evaluation of cokes made from South African coking coals and blends, as the coals

Executive HOWARD N EAVENSON, Chairman FREDERICK M. BECKET EDGAR RICKARD ERLE V. DAVELER WILLIAM WRAITH Finance HENRY KRUMB Chairman PAUL D. MERICA J V W REYNDERS H G MOULTON, Consultant Admissions JOHN M LOVEJOY, Chairman GEORGE D BARRON - DONALD M LIDDELL ELI T CONNER - WILLIAM W LYNCH

Jan 1, 1935

E. O'C. Acker ?86 'Truman H. Aldrich '78 Anson W. Allen '87 John H. Allen '84 W. Lawrence Austin ?82 W. S. Ayres '73 David Baker '86 'John Henry Banks '83 'Wm. de L. Benedict '88 'Robert Bentley '84 'Andrew A. Blair '75 'Walter M. Brodie Frederick Burbidge '87 'H. H. Burden '83 'Charles Butters '88 'R. C. Canby '82 Arthur C. Carson '84 'H. M. Chance '74 J. Parke Channing '84 Frank D. Chase '84 Clarence Raymond Claghorn '84 M. E. Clark '86 W. M. Claypool '84

Jan 1, 1937

By I. C. L. Webster, G. E. Ray

"The magnetite-rich Heff Cu-Au skarn lies in the Quesnel Terrane of south-central British Columbia, approximately 26 km northeast of Kamloops. The skarn formed close to the northern margin of the Heffley Creek pluton, and is hosted by Carnian (Late Triassic) Nicola Group limestone. The Late Triassic pluton (208.1 ± 6.1 Ma, U-Pb zircon) is a composite body of alkaline affinity that includes dioritic, gabbroic, and magnetite-rich clinopyroxene ± olivine ultramafic phases.The skarn has been discontinuously traced along strike for more than 1 km, and individual skarn layers are up to 40 m thick. The most common non-opaque minerals are garnet, clinopyroxene, and carbonate, with lesser epidote, biotite, amphibole, and chlorite. Apatite locally forms more than 3 vol.% of the rock. Compositionally the garnet is a low Mn grandite averaging Pyr3-Grs41-Adr56 mol.%.Mineralization consists of lenses, up to 10 m thick, containing massive low Ti-V magnetite (avg. 0.03 wt.% TiO2 and 0.01 wt.% V2O3). Magnetite is associated with up to 20 vol.% pyrrhotite, lesser pyrite, and trace chalcopyrite and gold. There is a moderate enrichment in rare earth elements with assays up to 570 ppm La and 490 ppm Ce. The Heffley Creek pluton belongs to a suite of Late Triassic-Early Jurassic alkalic and Fe-rich intrusions that are widely developed throughout the Quesnel Terrane. Some host Cu-Au porphyry mineralization, but less commonly they are associated with magnetite-apatite veins such as at the Glen Iron mine and Magnet occurrence, or with magnetite-rich Cu-Au skarns such as the Heff occurrence.Although abundant low Ti magnetite is present in the Heff skarn, it does not appear to represent iron oxide copper- gold (IOCG) type mineralization. Microprobe analyses completed during this study show that low Ti magnetite is not a unique feature of IOCG deposits, but also characterizes some other types of hydrothermal mineralization. In British Columbia these include some alkalic porphyry-related magnetite-apatite veins in the Quesnel Terrane, as well as the well-known Wrangellia-hosted Fe skarn deposits. Thus, low Ti magnetite appears to be a common signature of many magmatic hydrothermal deposits in contrast to the high Ti primary magnetite found in most igneous rocks."

Jan 1, 2007

By Christopher J. Bise

INTRODUCTION Perhaps no other area reflects the art as well as the science of mining-engineering design quite as obviously as strata control. This is due to the fact that the design of mining structures is difficult because rock rarely behaves as a homogeneous medium. Further, the original state of stress and the mechanical proper- ties of the in situ rock are somewhat difficult to deter- mine. However, those aspects that are hard to treat mathematically can and have been successfully approached by model studies and in situ measurements. This chapter will review the fundamentals of strata- control design and will provide examples of applications to typical problems. STRESSES AROUND MINE OPENINGS A close examination of Fig. 1 reveals that, prior to mining, stresses exerted in the overburden can be in both the horizontal and vertical directions. In the example problems of Chapter 1, it was shown that the maximum compressive stress (uv) at any depth (D) below the surface can be approximated as follows: [ ] Further, the lateral stress (oh) is related to the maxi- mum compressive stress by utilizing Poisson's ratio (µ): [ ] As soon as a heading is driven, such as the one in the coal seam shown in Fig. 2, the stresses shift to the sides, or abutments, of the opening. Clearly, these compressive stresses around the opening exceed the pre-mining conditions. The problem of designing an underground opening, therefore, is to determine the maximum stresses and to note if they exceed the ultimate strength of the rock. Table 1 lists several rocks with their corresponding strengths. Figs. 3 to 5 depict the boundary stress concentrations around circular, ovaloidal, and rectangular openings. M is the ratio of the horizontal stress to the vertical stress, or as defined above, p/(l-p). This text will concentrate on the state of stress represented by M = 0.33, which corresponds to the condition of no lateral deformation in a rock having a Poisson's ratio of 0.25. For an ovaloid with an opening width-to-height ratio of 2.0 (Fig. 4), the boundary stress concentrations are at their maximum values at the comers of the opening, with values in excess of 3.0. If the depth of overburden is 500 ft, then the maximum boundary tangential stress ([ ] ) of the ovaloid is: [ ]

Jan 1, 1986

By R. L. Grayson, G. Mishra

Introduction Many old underground mines have been developed extensively into reserves requiring large numbers of workers to maintain and rehabilitate lengthy track haulages and/or belt conveyor systems. Often, large numbers of fans are required to maintain minimal airflows in areas where poor conditions and huge leakages prevail. Problems that beset one such coal mine and projects undertaken to restore it to potential profitability are presented. Brief history The Vesta No. 5 mine was opened in 1908. Drift development began near the Monongahela River, which was the primary avenue for shipping metallurgical coal to coke ovens in Pittsburgh. Over the ensuing 75 years, the main haulage entries were developed a distance of 16.1 km (10 miles) to the present mining area. Track haulage was used to transport coal from working sections to the river loading facility. As haulage distance and areas of high water influx increased, management decided, in 1973, to construct an underground conveyor belt system. The project spanned 36 months and culminated in a system of 22 flights, ranging from 205.7 to 1767.8 m (675 to 5800 ft) long. Six intake and eight return entries were developed as the mains were advanced. These entries provided adequate ventilating capacity over the years, but as extensive gob areas progressed and roof conditions deteriorated, the entries became inaccessible and more resistant to airflow. By 1981, five fans were being used to circulate 943.9 m3/s (555.5 cfm) of air through the mine. An average of 6440 t (7100 st) were produced from 20.8 continuous miner unit shifts and 2.7 longwall unit shifts each day in 1981. The ratio of support to production personnel was 1.94. To maintain the outby 16.1 km (10 miles) of haulage, 110 personnel were used. The average cost per ton was $43.54 ($39.58 per st), resulting in excessive bottom-line losses. The mine possessed a rather promising reserve, mostly in lease, of good quality, medium-sulfur metallurgical coal, but it also had a number of formidable operating problems. Operating problems Contributing factors toward relatively low productivity and high costs were the conditions of the track haulage, conveyor, and ventilation systems. Productivity was affected greatly by downtime associated with the conveyor system, which lacked adequate access for making splices or other repairs. Many areas within the conveyor belt entry itself required more

Jan 3, 1987

By Dan Alexander, Ivan Crnkovic, Sarah Schwarz

"Several plant studies have been conducted at the Zinifex Century Mine over the past 5 years. For each of the plant studies, surveys of the zinc primary circuit were conducted, characterisation measurements for each operating cell taken, and the floatability components of the ore mapped to develop a model of the circuit. The floatability components and machine parameters were then used to simulate changes in circuit configuration and operating conditions, enabling further potential improvements in circuit performance to be identified.This paper reviews the current flotation practice at the mine, as well as various circuit changes and developments implemented by site. The application of the methodology and the evolution of the standard procedure are also discussed.INTRODUCTIONBase metals were first discovered in the Lawn Hill region of Queensland’s North West in 1869 and the first mining lease in this area was pegged in 1887 over what is now known as the Silver King vein deposit. The area encompassing these veins, which includes the Century deposit, was proclaimed the Burketown Mining Field in 1899. Exploration titles were granted to CRA in 1987, precisely 100 years after the initial discovery at Silver King, hence the name ‘Century’.Century was commissioned by Pasminco in late 1999, with the mine becoming fully operational on 1 March 2000 and the official opening taking place in April. Pasminco had bought the Century and Dugald River deposits from CRA in 1997 for A$345 million, and invested some A$710 million in bringing Century into production. Today, Century is the world’s second-largest zinc mine in terms of its output, having reached full capacity in 2003.The Zinifex Century Limited (“ZCL”) lead-zinc-silver deposit is located approximately 250 kms north-north west of Mt Isa in north-western Queensland. The deposit is a large flat-lying ore body covering an area of some 1.2 by 1.4 km, the 45 m-thick stratabound sulphide orebody is hosted within an 850 m-thick sequence of Proterozoic siltstone, shales and sandstone, amenable to open-cut mining methods. The mine has ramped up to full production delivering five and a half million tonnes of ore per annum, yielding approximately 900,000 tonnes of zinc concentrate and 90,000 tonnes of lead concentrate. Mine life is expected to be about 8 years"

Jan 1, 2008

By Hiroyuki Tamaru, Futoshi Kusumoto, Kosuke Tanimura, Atsuyuki Kimura

"SUMMARY The Hachinoshiri tunnel is a 2,469 m long two-lane road tunnel on the Chubu-Odan Expressway. Since more than 70% of the total tunnel length consists of low strength rock masses the competence factor of which is less than 1.0, it is expected that a self standing span cannot be achievable because of no ground arch to be established. The tunnel was thus designed with tunnel support patterns forming multi-arc early ring closure that can develop the load-carrying capacity estimated from the exerted earth pressure, Po, which can result from the competence factor, cf. The tunnel support pattern was selected based on the competence factor at tunnel face, then the subsequent tunnel support patterns were determined with reference to the monitoring data. The tunnel construction adopted full-face excavation accompanied by early ring closure using a 330 kW road header. The proposed construction methods at this tunnel site have turned out ensuring tunnelling in massive squeezing rock, so that they can be verified as reasonable tunnelling technologies, clarifying the relationship between the characteristics of the mechanical behaviour of the tunnel and its design parameters. INTRODUCTION The presence of highly weathered mudstone having the competence factor less than 1.0 where the tunnel became unstable with the early ring closure having the distance to early ring closure, Lf, of 6 meters, resulting in difficulty in maintaining the inner section of the tunnel, prompted a substantial amount of high strength double support and replacing of the tunnel support (Sato et al., 2015). Based on the mechanical characteristics learned from the construction, the mechanical stability of multi-arc tunnels was investigated using 3D FDM analyses with respect to the ratio of the radius for early ring closure, r3/r1, and the distance to early ring closure, Lf, so that tunnel support patterns for early ring closure can be established in terms of the competence factor, of, which was employed for a 725 m long section. Consequently the tunnel support patterns for early ring closure which can be determined from the competence factor became applicable to tunnel construction. The relationship between the characteristics of the mechanical behaviour of the tunnel and its mechanical parameters was also clarified. The early ring closure method adopted ensured tunnelling in massive squeezing rock, verifying its effectiveness in stabilizing the tunnel."

Jan 1, 2016

By Michael Moser, Brighty Dutta, David Dreisinger

"Molybdenum is widely used in the industries as an alloying element in the manufacture of steels and super alloys. The widespread use of the element encourages a growing need to develop extraction processes for recovery from both primary and secondary sources. The Solvay Group has developed a solvent extraction reagent named CYANEX®600 for the recovery of molybdenum from acidic solutions originating from both low grade (e.g. copper solvent extraction raffinates, smelter dusts and slags) and high grade (molybdenite) molybdenum sources. CYANEX®600 is a purified form of CYANEX®272 with the active component bis(2,4,4- trimethylpentyl)phosphinic acid. The extraction behavior of CYANEX®600 towards molybdenum in sulphuric acid medium was evaluated. The uptake of molybdenum increased with the increase in concentration of CYANEX®600. The approximate stoichiometry of the extracted complex was: Mo: CYANEX®600 = 1:2. Efficient back-extraction could be achieved with 0.1 M disodium EDTA and 0.1 M (NH4)2CO3, thus providing the option of recyclability of the extractant. The detailed back-extraction studies are currently under progress to find out the possible option for complete transfer of Mo(VI) to the aqueous phase so that high purity Mo salts could be precipitated from the molybdate solutions.INTRODUCTION Molybdenum (Mo) is one of the world’s valuable metals and is highly sought after by researchers due to its versatile mechanical, thermal, chemical properties. Swedish scientist Carl Wilhelm Scheele demonstrated molybdenite to be a mineral sulfide in 1778. Peter Hjelm successfully isolated dark metal powder from the mineral sulphide and named it as ‘Molybdenum’ in 1782 (Gupta, 1992). Molybdenum is a refractory metal of silver gray appearance and is an important alloying element used in the manufacture of steels and super alloys. Molybdenum compounds are widely used in formulation of catalysts (Gupta, 1992). Molybdenum occurs in nature only in chemical combination with other elements. Several molybdenum bearing minerals have been identified, but the only one of commercial significance is molybdenite (MoS2). As molybdenum is a strategic metal for the industrial sector, its widespread use acts as a factor for depletion of the natural resources. Hence, there is growing need to develop extraction procedures for its recovery from both primary and secondary sources. United States, China, Chile, Canada, Peru, and Mexico are listed as the main producers by the U.S. Geological Survey, 2016 (U.S. Geological Survey, Mineral Commodity Summaries, January 2016)."

Jan 1, 2017

By D. Gu, R. G. Anthony, C. V. Philip, M. E. Huckman

"In this paper, we review the equations needed to construct an ion exchange column model. These models contain three major components: 1) a differential material balance around the column, 2) a differential material balance around an ion exchange particle, and 3) a constitutive equation for diffusive flux. We then examine the ability of a fairly simple mathematical model to predict ion exchange column experimental data. This model contains the traditional absorption theory equations and is used to predict the behavior of a column filled with a hydrous crystalline silico-titanate and which is removing trace amounts of Cs+ from a high pH, complex electrolyte solution. Finally, we propose replacing Fick's law with the Nemst-Planck equation as the constitutive equation for diffusive flux. We use a NemstPlanck type model to predict batch experimental data, then evaluate it by comparing it's ability to predict experimental data to that of a more traditional Fick's law type model. The data from this comparison suggests that the Nemst-Planck type model yields better results.IntroductionMillions of gallons of strongly basic, radioactive waste are currently stored in underground storage tanks at Hanford and Savannah River and several locations within the United States. Due to the limited life expectancy of the tanks, there is a pressing need to remediate these wastes. This need has prompted the recent development of new ion-exchange materials for removing radioactive elements from aqueous solutions. A novel material called TAM-5 has been developed jointly by Texas A&M University and Sandia National Laboratories. TAM- 5, a crystalline silico-titanate, is highly selective for Cs + in the complex electrolytic solutions typical of these wastes. It is stable in very basic solutions and in radioactive environments [1, 2 ,3]. In laboratory testing to determine the selectivity of various materials, TAM-5 was superior to other candidate ion exchangers [4, 5]. The selectivities were determined by measuring a distribution coefficient, Kd, which is defined as the concentration, gig or mmols/g of Cs in the solid phase divided by the concentration, mg/L or mmols/mL, of Cs in the liquid phase. The units of Kd are ml/g or L/kg. Table 1 shows Kd values for TAM-5 compared to 9 other candidate materials in a typical waste solution [6]."

Jan 1, 1996

By Christopher E. Dianora, Matthew Sorrell, Eileen Test

"Constructed between 1939 and 1945, the Rondout-West Branch Tunnel (RWBT) portion of the Delaware Aqueduct is 45 miles (72 km) long. The New York City Department of Environmental Protection (NYCDEP) first identified suspected areas of leakage from the RWBT in the 1990s. A $1 billion project is underway to mitigate the leakage through the construction of a bypass tunnel around the area of most severe deterioration, a 2.5-mile (4 km) section located beneath the Hudson River. Two shafts, 800 and 900 feet (244 and 274 m) below the ground surface, are being sunk to support the tunnel construction. The deep shafts will be subjected to high groundwater pressures and have been designed to accommodate multiple tunnel portals. This paper discusses the engineering decisions and analyses leading up to the final design. INTRODUCTION Project Background The New York City Department of Environmental Protection (NYCDEP) provides over 1 billion gallons (3.7 billion liters) of water to over 9 million residents of the City of New York and surrounding communities each day. This water is stored in upstate reservoirs and conveyed to the City via a system of aqueducts and tunnels. Water is delivered to the City almost completely by gravity and consists of three systems: the Croton, the Catskill, and the Delaware. The Delaware System is the largest of the three and supplies between 50% and 70% of the City’s daily water supply. In the 1990s, the NYCDEP discovered leaks in the central 45-mile-long (72 km) portion of the Delaware Aqueduct known as the Rondout-West Branch Tunnel (RWBT). The leaks release between 15 and 35 million gallons per day (MGD) (132.5 million liters per day) and are concentrated in two areas near Wawarsing and Roseton, New York. As part of the Rondout-West Branch Bypass Tunnel and Wawarsing Repairs Project (Bypass Tunnel Project), the leaks in the Wawarsing area will be repaired by grouting from inside the existing RWBT, and the leaks in the Roseton area will be mitigated by constructing a bypass tunnel parallel to and 1,750 feet (533 m) north of the leaking portion of the existing RWBT. Two deep shafts will be excavated on either side of the Hudson River: Shaft 5B in Newburgh and Shaft 6B in the Town of Wappinger."

Jan 1, 2016

By Martin Junker

This Notwithstanding all the debates on climate change and renewable energy, coal remains the world's no.1 primary energy source even in the 21st century. In 2011 alone, total output rose from 420 million t to around 6.6 billion t. Similar growth rates are also expected for the next few years. In order to meet demand, new seams are being tapped and others mined far more extensively. In most cases, this means a move from open-cast to deep mining, and extracting coal from ever greater depths. At the same time, extraction processes such as the room and pillar method are being replaced by longwall mining, which significantly increases seam yield and therefore output. The depth of extraction with new, modified mining methods is increasing rapidly around the world - in most countries, much more rapidly than the corresponding expertise. A trade in mining expertise has therefore developed alongside the coal trade proper. Germany relies on coal imports, but can export mining industry services and advanced mining equipment. Geological conditions in the Ruhr and Saar coalfields forced Germany to drive forward development of deep working for hard coal years ago. For 30 years now, Germany has been mining hard coal at depths of over 1,000 m, and excavation is now at depths up to 1,600 m. Deeper extraction and more extensive excavation posed new challenges for both miners and equipment. Intelligent logistics, modern roadway development methods and longwall mining with ploughs or shearers combined with optimised process coordination offered pioneering solutions - solutions to enable costeffective coal extraction in the long term without compromising safety. Countries such as Russia, the Ukraine, China and the Czech Republic, where extraction is now reaching similar depths, can now benefit from this experience: they do not need to go through the same process of trial and development, but can draw on Germany's findings and ultimately save on costs. Now, Germany is offering this knowledge and experience on the international market on a client- and deposit-specific basis to continue providing mining countries with safe and cost-effective coal mining solutions and secure the energy supply. The presentation gives a number of examples of how international mining companies have used expertise from Germany to better meet the challenges of deep mining.

Aug 1, 2013

By Kimio Itagaki, Torn Matsumoto, Makoto Ohtsuka, Kentarou Harada

"The structure and hydrogen-absorption properties of LaNis thin films were investigated in the present study. LaNis thin films were prepared by a sputtering method with a radio-frequency power, Ws, of 50 -400 Wat substrate temperature, Ts, of323 -573 K. Nickel content of the films decreased with increasing Ws, however, there was no dependence on Ts. At low Ws and low Ts, amorphous films were deposited. In other conditions, the films were polycrystalline. The films prepared at high Ws and low Ts, or at low Ws and high Ts, exhibited only a (h k · 0) peak in the X-ray diffraction profiles. This indicates that the c-axis of crystalline films was oriented parallel to the substrate plane. At high Ws and high Ts, the films with columnar structure were obtained. Hydrogen-absorption capacity of the thin films increased with increasing Ws. The pressure plateau did not appear on the pressure-composition isotherms. The films deposited at high Ts represented a higher equilibrium pressure.IntroductionThin films of hydrogen storage alloys can be applied for an electrode of rechargeable battery or fuel cell, a filter for H1 purification and various kind of sensor. Some studies on the preparation of the thin films of LaNi5 intermetallic compound, which is a typical hydrogen storage material, have been done hitherto. Adachi et al. , 2) and Sakaguchi et al. 3) prepared amorphous and crystalline LaNi5 thin films using flash evaporation of alloy powders and measured the pressure-composition isotherm for the prepared film utilizing a quartz crystal mass monitoring (QC:MM) method. It was found that the prepared amorphous film absorbed hydrogen without disintegrating into the powders and the hydrogen-absorption capacity of the films increased with the degree of crystallinity. Sakai et al. 4) reported that the maximum hydrogen storage capacities of amorphous and crystalline LaNi5 thin films, fabricated by a RF sputtering method using combination targets, were about 40 mA h g-1 (LaNisHo.7) and 160 mA h g-1 (LaNi5H2.6), respectively. Uchida et al. 5, 6) prepared LaNi5 thin films using flash evaporation of alloy powders and measured volumetrically the pressure-composition isotherm for the prepared film. It was reported that the hydrogen-absorption capacity of the films changed by the annealing treatment."

Jan 1, 1998