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By B. P. Boisen

INTRODUCTION Application Measurement of rock deformation is essential to stability and behavioral monitoring of underground openings. The difficulties of obtaining adequate sample distribution, density, and orientation can frequently be overcome by the use of single and multiple position borehole extensometers. In its simplest form, a borehole extensometer consists of an instrument head, usually anchored in the collar of a borehole, and a number of in-hole anchors, each of which is fixed in place at a known distance from the collar. As rock adjacent to the borehole deforms, the component parallel to the borehole direction causes a distance change between adjacent in-hole anchors, as well as a distance change between individual anchors and the instrument head. These changes are measured by means of a mechanical link (usually wire or rod) extending to the instrument head from each anchor. The resulting data can be used to compute the distribution, magnitude, rate, and acceleration of deformation in the material intersected by the extensometer's borehole. Wire Vs. Rod Extensometers Since the early 1960s myriad varieties of borehole extensometers have become available through the combination of various special features of two basic instruments, the wire and rod versions of borehole extensometers. Recent publications have recommended shallow depth limitations for wire extensometers, and favor rod extensometers for a broad range of borehole inclinations and depths. There are three main reasons for using a deep wire system: its cost effectiveness; the fact that tremendous depth means tremendous rod weight which will cause the rod to buckle or rub against the borehole or casing wall; and users' attitude and proper understanding of how to install and apply a wire-type extensometer. Here experience plays a big role. Over the years, many have learned the hard way that untensioned rod extensometers are simple and inherently inexpensive, but limited to vertical up holes where sensitivity is not influenced by friction between the rod elements and the borehole wall or protective conduit, the inertia of the rod elements, or the direction of the movement to be measured. In deep, horizontal, or inclined holes, and in instances where compressive as well as tensile strains or displacements are anticipated, tensioned rod or wire systems are preferred. System Components Borehole extensometers are made up of at least the following five principal components: the instrument head; the collar anchor assembly; the downhole system of anchors and rods or wires; the signal cable if the head is equipped for electronic readout; and the readout apparatus, whether mechanical or electronic. These components are shown in Fig. 1. Borehole Requirements Most extensometers are designed for installation in nominal 75 to 57-mm (3.0 to 2.25-in.) diam drill holes. Either a diamond drill or a percussion drill hole is satisfactory, provided the variation in drill-hole diameter is within the tolerance of the standard anchor design. The basic system may be modified for other hole diameters by changing the dimensions of the collar anchor and downhole anchor assemblies. Some instrument heads and collar anchor assemblies require enlarged borehole collars, though usually not over 150 mm (6 in.) for a depth of 1 m (3 ft). It is desirable, although not essential, that holes be free-draining. Holes must be clean and free of debris before installation. Depth/Orientation of Extensometer Boreholes Extensometers are designed to measure axial strain components in the instrumented drill hole. As the drill hole is axially deformed, the distances between individual downhole anchors and the instrument head change, as do the distances between adjacent anchors. The distance changes measured by a multiple position borehole extensometer, for example, are reduced to functions of time, hole depth, anchor position, or other parameters of significance in terms of the structure under investigation.

Jan 1, 1982

By J. C. Agarwal

Major scientific discoveries cannot be programmed. New or frontier technology, however, can and should be planned. It is the vital force underlying the business strategy of a vigorous and profitable mining and metals company. Generally, a radical departure from existing technology is required to 1) Enhance the competitive value of ore reserves, i.e., to turn low grade or marginal reserves into competitive reserves. 2) To increase yield and decrease costs. 3) To decrease investment. 4) To increase mine life. 5) To improve product quality to meet market demands. The very definition of frontier technology implies that it is so different that the old plant, if already in use, will not be directly suitable to use the new technology. Therefore, a priori, we know that the commercialization of frontier technology would very likely involve new investment. The recommended technique of economic and technical analyses, I propose to call "Success Analysis.'' Simply stated, it forces us, by step- by-step procedure, to ask "what is needed to succeed in this particular project." Therefore, the first step is to identify what is meant by "success." The definition of "success" requires an agreement with top management or sponsors of the project as to what constitutes "success" in their minds as contrast to the technical people perception of "success." In general, top management would define success as meeting: 1) The technical objectives, i.e., technology being developed would produce the desired product at the acceptable quality. 2) Financial objectives, i.e., when commercialized, the project's capital cost, operating costs and return on investment are better than the projects they would normally fund. The return has to be better because new technology is inherently riskier. 3) Marketing objectives in that the project will produce product of acceptable quality in timely manner, i.e., scale-up to commercial plant would not cause undue delays in start-up and that large fix-up expenditure would not be required to protect the projected return on the investment. Many projects are started in the laboratory to develop "Frontier Technology" in mineral processing. Most of these efforts are never commercialized. There are many causes of these failures. The major causes are listed below: 1) Inadequate communication with the sponsors of the project so that project objectives are defined and agreed to by all. 2) Ill-defined technical problems. 3) Incomplete development of technology. This has two major components: a. Incorrect scale-up and engineering. b. Inadequate start-up and maintenance planning. 4) Incomplete and over-optimistic process economics. 5) Lack of understanding of the future behavior of the marketplace. 6) Failure to complete R&D effort by the time it is needed. 7) Incorrect projections of human resources required; that is, technical, managerial and operating. 8) Lack of realistic forecast of the expenditures required for commercialization of the R&D effort. These pitfalls can be avoided by comprehensive planning and the use of success analysis technique. The major elements in "Success Analysis" are: 1) Identification of the "need" for new technology. 2) Unbiased and complete scale-up engineering. 3) Acceptable economic feasibility with enough margin of error in the sale prices and operating costs to justify the inherent risks in frontier technology. 4) Timing for availability of investment capital and markets. Many failures in the development of frontier technology are preordained because the initial selection of the research project has been without adequate attention to the needs of the industry. Many very innovative hydrometallurgical processes for recovery of copper from concentrates were developed in the past twenty years. They have not adapted except to a very minor extent because they did not satisfy the success analysis criteria. The copper industry in North America needed a process to compete against high grade ores and to avoid

Jan 1, 1985

Although there was a significant improvement in the world economy, recovery in the iron and steel industry followed its traditional pattern and lagged the general economic recovery in 1983. World crude steel production increased from 645 Mt (711 million st) in 1982 to 664 Mt (732 million st) in 1983, an increase of less than 3%. Domestic iron and steel production showed a modest increase over 1982. Weekly raw steel pro¬duction levels ranged between 1.5 and 1.75 Mt (1.6 and 1.9 million st) for the last 10 months of 1983, compared with less than 1.25 Mt (1.4 million st) for the last half of 1982. The peak was 2.5 Mt (2.7 million per st) per week in 1978-79. This how level of steel operations kept most pelletizing and sintering operations at curtailed production levels. Extended shut-downs of pellet plants were also common. Pressure by management to reduce operating costs has continued with an increasing emphasis on maintaining quality and efficient unit productivity. With the extended decline, there have been significant personnel reductions in both operating and headquarters staffs. Many companies are anticipating a long-term how growth period in the iron and steel sector. Pelletizing Domestic pellet production showed a slight increase last year to about 34 Mt (38 million st), less than 50% of total capacity. What is indicative of the inevitable is the announced intention to chose down facilities, such as US Steel Corp.'s Atlantic City ore operations pellet plant at Lander, WY, and the recent aborted move by US Steel to purchase National Steel with the apparent intent to chose the National pellet plant in Keewatin, MN. There is excess domestic pellet capacity that will not be needed. The question is what form attrition will take. South American pellets can be brought into the Chicago area at a cost competitive with Great Lakes plants. This raises the possibility that imported pellets can, in fact, displace domestic pellets at almost any US steel plant. The need to re duce total production costs at domestic pelletizing plants is crucial. Efforts to reduce fuel costs for pelletizing have focused on re placing natural gas or fuel oil with coal. More recently, petroleum coke has become available at an attractive cost, but systems must cope with its higher sulfur con tent, which limits its potential The direct firing of grate-kiln systems with pulverized coal hay been fully proven. The use of solic fuel in the straight-grate process however, is still being developer and no straight-grate system i running on solid fuel alone. The further development o coal-water slurry systems may provide a suitable fuel for both systems. It would appear, how ever, that the same ash quality restriction would hold for coal water systems as for direct-firin systems. So, the choice is based of the relative cost and convenience of a wet-grinding system with additives, compared with a dry-grind direct-fired unit. The high-temperature (two stage) heat recuperation system installed at US Steel's Minnta plant (Step III) has provided the fuel savings anticipated, any modification of additional in durating lines is planned. Other operators are hooking at similar heat recuperation improvement. Another area of potential cost savings has been in the reduction of bentonite requirements. A CCI's Tihden plant, major improvements have been made in this area. Progressive bentonit reductions were achieved over several years from the 8- to 9-kg (16- to 18-lb per st) range to the 5-k (12-1b) level by progressive moisture reductions and chose attention to operating practice. Further reductions were achieve by raising bentonite specifictions. In this instance, the wate plate absorption test specifia

Jan 5, 1984

By Feng Cai

"As a new method to support gate roadways, rock bolting has been widely used in the big coal enterprises for its superiority. The disturbed stress field of surrounding rock and anchor pre-stress field were divided into 3 areas in consideration of s1, s3. The surrounding rock failure form and characteristic were analyzed in combination with parameter f, the failure mechanism was explained using the slip line theory, furthermore, the rock bolting mechanism and its arrangement were studied. After roadway excavation, the tensile stress area appeared at the surface of the surrounding rock. The failure type changes from shallow tension-shear form into deep compression-shear form and the failure shape is an “incense burner”, surrounding by high risk failure area in butterfly shape. The maximum principle stress in roof appears discontinuous distribution because of layers separation. The failure type of surrounding rock is intensity-stress environment-weak plane control model. The internal reason of roadway macroscopic failure is the displacement between the surface plastic area and the deep elastic region along the tangential direction of slip line. In condition of rock bolt support, there is one-way tensile stress area in small range at the surface of surrounding rock, inside is the bidirectional compressive zone in uniform distribution form, stress environment is improved by the superposition of anchor pre-stress and disturbed stress field, anchor rock bolt arrangement and pre-tightening force distribution is decided according to the distribution of f parameter. Finally, supporting method and parameter reasonability was proved by the roadway deformation monitoring, damage depth drilling peep test in the 8101 transportation roadway in a mine.Introduction The stability of roadway support is very important to the safety and sustainable development of coal mine. In recent years, the focus of China's coal mining industry began to move to West China, the thick-extra thick coal seam is developed in the West. The latest technology and equipment have been put in to service, the occurrence condition of thick coal seam and the increasing of productivity lead to the cross section of the roadway increasing continually, hence, it is more difficult to control the surrounding rock. At present, the most common way to support the roadway is rock bolt (cable) support (Wang 2007); Scholars have been deeply studied on the mechanism of rock bolt. It is reported (Wang et al 2012, Gao 2007, Zhang and Yuan 2006) that the influence of rock bolt on physical and mechanical parameters of coal and rock mass in the anchorage zone, It is believed that the rock bolt can improve the peak strength of surrounding rock, especially the post peak strength, and the failure mode of rock mass under the rock bolt support is analyzed. It is reported (Wang et al 2014, Xiao et al 2013, Li et al 2007, Yan et al 2012), according to different geological conditions, the failure mechanism of surrounding rock and supporting technology of the whole coal roadway were studied. Focus on the failure mechanism of roadway surrounding rock, different forms of high strength support methods were proposed, which achieved good results. It is reported (Wang et al 2012a, Zhang et al 2011, Wang et al 2012b, Kang et al 2010, Zhang et al 2010, Kang et al 2007) that the supporting mechanism of pre-stressed anchor was analyzed from the point of the distribution of the maximum principal stress in the pre-stressed anchor field, and the influence of the rock bolt support parameters and surrounding rock characteristics on the supporting effects were studied. Based on the latest research results, this paper analyzed the failure mechanism of surrounding rock of coal roadway, proposed the high strength anchor net cable combined supporting technology based on the feature of stress zoning distribution of surrounding rock, to provide reference for the coal roadway rock bolt (cable) support technology."

Jan 1, 2017

By James C. Irvine

Pea Ridge Iron Ore Co., previously Meramec Mining Co., a joint venture by Bethlehem Steel Corp. and St. Joe Minerals Corp., mines and pelletizes iron ore at the Pea Ridge mine. The Pea Ridge property, now wholly owned by St. Joe, is located near Sullivan, MO, about 112 km (70 miles) southwest of St. Louis. The ore body was delineated in the mid-1950s by St. Joe during a lead exploration program. The first test holes drilled on the Pea Ridge magnetic anomaly revealed the presence of a large magnetite deposit; further drill¬ing in 1956 and 1957 confirmed that the ore body was minable. Meramec Mining Co. was incorporated in 1957 and shaft sinking began late in the year. Pro¬duction commenced in April 1964. The ore body is overlain by about 396 m (1300 ft) of flat-bedded sediments. It is tabular, about 792 m (2600 ft) long, and up to 182 m (600 ft) thick. It dips about 1.39 rad (80°) and is of unknown depth. The ore is mainly high grade magnetite with small zones of specular hematite. The wall rock is a Precambrian rhyolite porphyry. The mine was initially started with five major levels on 45-m (150-ft) intervals. Crosscuts were driven across the ore body on 58-m (190-ft) centers. Banks of stopes were mined between the crosscuts by a modi¬fied shrinkage stoping method (Fig. 1). This was done by undercutting a 12 m (40 ft) wide by 45-m (150-ft) long block and blasting horizontal "lifts" drilled on 1.5-m (5-ft) intervals. A 20-m (65-ft) sill was left between levels which contained the slushing drift, fin¬gers, crown pillar, and adequate thickness for support. By 1971 a "lattice" of pillars had been left and mining had progressed to the point that an orderly pillar re¬covery program was necessary (Fig. 2). INITIAL PILLAR RECOVERY PROGRAM The program was started in the upper western por¬tion of the ore body, furthest from the shafts. Due to the fact that the ore body narrows in the western extremity, the stope orientation was changed to mini¬mize required development. This left pillars 18 to 24 m (60 to 80 ft) wide and up to 79 m (260 ft) in length. The area selected for the start of the program had been developed between the uppermost level at 419 m (1375 ft) below surface and the 510-m (1675-ft) level. Due to the 91-m (300-ft) difference in depth, sublevels were driven on 15-m (50-ft) intervals and the ore was taken by rather classical sublevel stoping methods (Fig. 3). This left a structure with accesses compatible with drilling with 114-mm (41/2-in.) bore drifters mounted on columns. Most holes were drilled 63.5 mm (21h in.) diam and reamed to 100 or 127 mm (4 or 5 in.). Practical depth capability was about 21 m (70 ft) for holes above horizontal and 13 to 15 m (45 to 50 ft) for holes below horizontal. The interdependence of these long pillars necessitated that several of them be blasted simultaneously, making large blasts the only practical approach. Due to the required length of such loading cam¬paigns, it was necessary to select an explosive which could stand in the hole in an underground mine for a 4 to 6 week period. After some consideration a pump¬ able water gel was selected for both uphole and down¬hole loading. This afforded a high velocity, high density explosive which could be handled in large quantities. Loading the downholes was a fairly straightforward process. The holes were lined with an 8 mil polybutylene plastic sleeve and water gel was pumped into them with a double diaphragm pump mounted on the bottom of a stainless steel tub with a 81-kg (180-1b) capacity. Potential leakage into cracked areas of the pillar was

Jan 1, 1982

By W. M. Ma, Daniel W. H. Su, K. Centofanti, Yi Luo, W. L. Zhong, Syd S. Peng

7.1 INTRODUCTION Subsidence measurements include surveys of surface and sub- surface movements and measurements and monitoring of surface structural damages. In this chapter the most commonly employed methods are discussed. 7.2 MEASUREMENT OF SURFACE MOVEMENT Surface movements involve two basic components, subsidence and displacement. The other five components (i.e., slope, curvature, strain, twisting, and shearing) are derived from the two basic components. Since all measurements are geometrical, they can be measured by common surveying techniques. 7.2.1 MONUMENTS The first step is to select the types of monuments or surveying stations. Various types of monuments have been employed. The simplest one is a paint mark on the pavement or concrete slab, although this should be discouraged because it is easily affected by artificial or environmental factors. Others use steel or wooden rods, 2 to 3 in. in diameter or square by 2 to 5 ft long, driven all the way down, flush with the ground. Both types of-monuments have no provision for frozen ground heave and both are tampered with easily. A more elaborate type is shown in Fig. 7.1 (O'Rourke et al., 1982). A rebar or iron pipe, 2 in. in diameter by 6 ft long is driven down to nearly flush with ground surface. The top 4-ft section of the hole around the pipe is enlarged to 1 ft in diameter and grouted with concrete. The most sophisticated type of monument is shown in Fig. 7.2 (Conroy and Gyannaty, 1983a). It consists of two hollow steel pipes. The inner I-in. pipe, 10 ft. long is enclosed by a 2-in. pipe at the upper 5-ft. interval. The gap between pipes is filled with foam to isolate the inner pipe from external environmental effects. The bottom of the inner pipe can either remain open or be sealed with a cap, depending on whether the ground is dry or wet. 7.2.2 MONUMENT ARRANGEMENT AND SURVEY LINES The monuments are installed strategically to form survey lines. Three basic types of monument arrangements are shown in Fig. 7.3. Figs. 7.3a and 7.3b are for single and multiple panels respectively; Fig. 7.3~ is mainly for research purposes with multiple survey lines both in longitudinal and transverse directions. This is extremely useful for defining the distributions of ground deformation, as shown in Figs. 2.10 and 2.11. Basically there are two types of survey lines: the whole basin survey line which extends across the panel width or length (e.g., lines R, R, in Fig. 7.3 a&b), and the half basin survey line which covers approximately half of the panel width or length (e.g., lines R, F in Fig. 7.3 &b). Survey lines are generally straight and linear. They should be located on the major crosssections of the subsidence basin. Survey lines must extend beyond the edge of the opening for a sufficient distance so that the edges of the subsidence basin can be determined accurately. The minimum distance from the edge of the opening must be [ ] where h = mining depth, [ ] = angle of draw, and [ ] = 5 to 7 times monument spacing for whole basin surveys, or 8 to 10 times the monument spacing for half basin surveys. There must be at least two controlling monuments at each end of each survey line. Controlling monuments must be located beyond the influence of underground mining. The spacing of the monuments ranges from 15 to 200 ft. The UK National Coal Board (NCB, 1975) recommended a spacing equivalent to 0.20 of the mining depth. Peng and Geng (1982) recommended that the maximum monument spacing be dependent on mining depths as follows:

Jan 1, 1992

By T. N. Moore, E. M. De Souza

J. Dasher Regarding the March 1991 ME technical paper by De Souza and Moore: For more than a decade since my February 1981 article on how to use modern metric, which SME-AIME had decided to do, I have monthly pointed out metric errors to the editors. In part, I do this because there has been no action to allow editors to fix figures and tables or to allow them to require authors to do so. The latest resulting atrocity provokes this discussion of vibrating drill handle units being stated in decibels. Reply by T. Moore We have read the discussion of our paper by Mr. Dasher. Our reaction is one of surprise and incredulity. It would seem that Mr. Dasher takes exception to the use of the decibel scale to present vibration acceleration data, and the use of hertz as the unit for frequency. The basis for his objection to the decibel appears to be that it has no dimensions (which somehow invalidates its use), that it is "non-metric" and, finally, that it is parochial (of limited or narrow scope). His objection to the use of the term hertz is not stated, but we will assume that it stands condemned as "non-metric" and parochial. Obviously we disagree with Mr. Dasher's views and will now outline our reasons. Although the decibel scale originates from transmission line theory and telephone engineering, it is also at present widely used, not only in the fields of electronic engineering and acoustics, but also in the area of vibration. The original definition of the decibel (dB) was based on power ratios: dB = 10 log 10(W/W0) where Wo is a reference power. However, as the power measured across a given impedance is related to the square of the force acting upon this impedance, Z, a more commonly used definition is: [2 dB = 10 logF /Z) = 20 log F/F 10\ F0 2 /Z(0)] where F and F0 are the r.m.s. values of the forces. Now, if the measurements are related to one and the same impedance, the decibel notation in the form of 20log10(X/Xo) may be used as a convenient relative magnitude scale for a variety of quantities. Thus, X may, for instance, be an r.m.s. displacement, velocity or acceleration. It is only required that XD always be a reference quantity of the same type as X. That is, when X represents an acceleration, then X0 represents a reference acceleration. This is the formulation used in our paper. This was not an arbitrary choice on our behalf but reflects standard practice as specified in the International Standard ISO 5349-1986(E) Mechanical Vibration - Guidelines for the Measurement and the Assessment of Human Despite the metric prefix, the decibel is a parochial expression of (l) the logarithmic ratio of the loudness of a sound to what is normally audible or (2) the logarithmic ratio of two power signals in radio or electronics. A decibel is not a unit, much less an SI, unit and has nothing whatsoever to do with the acceleration of drill handles. Stating that m/s2 (acceleration) is decibels is without reason. Whoever reviewed this material should not have allowed publication of figures of dB and H.[ ] Exposure to Hand-Transmitted Vibration. This was clearly stated in the "measurement protocol" section of our paper. This quantity is then referred to as the acceleration level and is expressed in dB. We may have inadvertently caused some confusion when we simply used the term acceleration to refer to acceleration level on our diagrams. At the time, we felt the use of dB or m/s2 would make the context clear to the reader. For any confusion this decision may have engendered, we apologize. Since the decibel expresses the ratio of two like quantities, it certainly has no dimensions. It is, however, common practice to treat "decibel" as a unit as, for example, in the sentence, "The acceleration level measured at the operator's hand was 160 dB." The expression of measured quantities in dimensionless form is not inherently unacceptable. In fact, in many areas of engineering it is standard practice (consider the use of Reynolds Number, Nusselt Number, etc.). The fact that the decibel is a dimensionless quantity makes the question of whether it is a SI unit nonsensical. However, it is valid to insist that the dimensional quantities used to obtain the decibel values be expressed in SI units. A careful reading of our paper will make it clear that the measured acceleration was, in fact, expressed in units of m/s2 as was the reference acceleration (l x 10-6 m/S2). These are the accepted derived SI units for acceleration. See, for example, the standard ASTM E380-89a Standard Practice for Use of the International System of Units (SI) (The Modernized Metric System). Concerning Mr. Dasher's implication that hertz (Hz) is an unacceptable unit of measure for frequency, we would again refer him to the standard ASTM E380-89a. Here, he will find (section 2.4.2) that hertz is an accepted "special name" for the derived SI units-1. This is in keeping with numerous other international standards including ISO 5349-1986(E) to which we referred in our paper. In conclusion, we agree with Mr. Dasher on the desirability of expressing measurements in modern SI units. But we would remind him that the standards that define the use of these units, and the accepted means of presenting measured data, are in a continual state of refinement. It is, therefore, incumbent upon him to keep abreast of these changes if he wishes to constructively critique the work of others.[ ]

Jan 1, 1992

By J. R. Schafler, J. L. Brinkman

"The US coal mining industry has suffered a recent decline in profits caused by a combination of decline in commodity demand and more restrictive government regulation, forcing coal mining companies to cease production in numerous mines. A 40% decrease in the total number of coal mines occurred from 1994 to 2000. The number of mines remained relatively constant until 2008 when a change of government sparked a series of new environmental regulations which have contributed to another 15% decrease from 2008 to 2016. In contrast, production has only decreased by 3%, increasing from 1994 to 2008, and then decreasing since 2008. However, this value is skewed by large increases in sub-bituminous coal production and decreases in other coal ranks. Several aspects assist in executive decisions on mine longevity and economic worth. Studies of open source mine data collected by mandated government surveys coupled with commodity history can be used to find relationships between mine closures and these elements, which include statistics on production tonnage, seam height, coal rank, etc. This paper looks at coal mining trends since 1994 by comparing mine, market, and regulation factors that may influence mine closure. INTRODUCTION The US coal mining industry is currently experiencing a decline in both demand and commodity price while facing implementation of more stringent air and water pollution regulations. Company profits have decreased significantly, forcing several major companies to declare bankruptcy. The industry was required to adapt to survive. Many of these companies have adjusted by abandoning mines; the last two decades have seen a drastic decrease in the number of mines. To discover the primary causes of this downturn, three major aspects must be studied: properties of abandoned mines, commodity history, and regulatory impact. TRENDS SINCE 1994 Each year, US coal mines are required to submit data on a wide variety of categories to the Mine Safety and Health Administration (MSHA) and the US Energy Information Administration (EIA). This data includes administrative records, production values, and information regarding the attributes of the mined material and the mine itself. Each year since 1996, EIA has released a report summarizing the data from two years prior. Additionally, MSHA maintains a database containing a collection of all the most current data. These resources are the primary sources for the following figures."

Jan 1, 2017

By Matthew Geary, Stephen J. Marino, Jordan A. Schreiner

"INTRODUCTION Project Overview The Southeast Collector (SeC) is a new 15km long 3.0m internal diameter wastewater tunnel designed by Hatch Mott MacDonald (HMM) that twins the existing 2.6m diameter York-Durham Sewage System (YDSS) tunnel. Both sewage tunnels are located north of the City of Toronto in Ontario, Canada. The SeC provides much needed conveyance capacity for the rapidly growing York Region population, supporting development until 2036 and increasing sewage system security (Hatch Mott MacDonald and AECOM 2011). Construction of the SeC will also permit flows of the twinned section of the YDSS to be bypassed in order to clean, inspect and carry out any essential maintenance within the YDSS tunnel. Although owned and operated by York Region, the SeC spans two regional municipalities beginning upstream from the intersection of Ninth Line and Rouge Bank Drive in the City of Markham in York Region to the downstream Finch Avenue and Valley Farm Road intersection in the City of Pickering in Durham Region. See Figure 1 for the SeC and YDSS alignment. The project includes 19 shafts and chambers with depths ranging from 8 to 48m in addition to seven facilities supporting the operation and maintenance of the sewer. The construction contract was awarded to Strabag Inc. in July 2011 for $291M CAD. Construction Management services were provided by a joint venture team of Hatch Mott MacDonald and AECOM. Existing York-Durham Sewage System The YDSS tunnel was completed in 1979 and was constructed using an open shielded TBM with ribs and lagging temporary support, followed by installation of a 380mm thick unreinforced cast-in-place (CIP) concrete final lining. The YDSS sewer system has numerous connections that accumulate flows from various urban areas within York Region. The section of the YDSS twinned by the SeC flows under gravity conditions and had been putting a strain on the municipality’s potential for population and development growth having reached its maximum capacity of 420 ML/d. According to the York Region’s Water and Wastewater Master Plan, the existing YDSS would not have been able to handle wet weather flows past 2016 (Genivar and XCG Consultants 2009). The dependability of this section of the YDSS is also of concern because it is the main and critical connection between the sewer systems upstream and the downstream Duffin Creek Water Pollution Control Plant (WPCP)."

Jan 1, 2016

By Lu Yang

In the last two decades, acid iron ore pellets have established their predominant position as the iron-bearing burden material for blast furnace ironmaking in North America. It is a general practice that pelletization plants are asked to produce pellets of certain chemical composition and in a narrow range of sizes. A minimum amount of fines created during handling and transportation is desirable. Such a requirement is often expressed indirectly by the minimum compression load causing a pellet to break. The behavior of pellets inside a blast furnace, of course, is very important for a smooth and efficient operation. Before its melting in the lower part of a blast furnace during this period of several hours the behavior of pellets depends on their properties after induration as well as conditions inside each furnace. Among properties of indurated pellets, the ability to maintain their physical integrity in the burden and to be reducible are most important. To maintain the integrity of pellets (i.e., minimal low temperature breakdown, swelling and softening) so that there will be adequate overall permeability for certain wind rate and local permeability for gas/solid contact for reduction. A higher reducibility will moderate the adverse effects of swelling and softening and contribute to a smooth operation and efficient use of heat and reductants in a blast furnace. An article devoted to the relationships between the quality of iron ore pellets and blast furnace practice has been published elsewhere(1); therefore, no further discussion is needed here. The mechanism of strengthening of iron ore pellets during firing has been the subject of investigation of many authors(2). There is no doubt that the bridging of the growing hematite grains, particularly starting with magnetite concentrates, is very significant in the production of strong pellets. Recently, there are many reports from blast furnace operators that they have noticed the adverse effect of LTB in their plants. It is obvious that the strength derived from bridged hematite grains may disappear when hematite is converted to magnetite with thorough breakdown of grain structures. It is suggested that some other strengthening mechanism may be required in order to improve the LTB index of acid pellets. The influence of porosity on both compression strength and reducibility follows regular patterns. It has been suggested(1) that the amount of liquid slag in pellets during firing determines to a large extent the porosity, hence, indirectly, strength and reducibility.

Jan 1, 1983

By F. Heivilin

"Health and Safety problems can make a mineral deposit or a portion of one, an operating plant, or finished product unprofitable or worthless. Therefore they need to be reviewed when reserves are established and yearly thereafter until the deposit is closed down. Health and Safety regulations serve as a guide, but I have always done more frequent checks, when adding new products with different specifications or during times of bad weather and crisis. Water quality and quantity are always in question. Health and Safety are quality of life issues which are at or near the top of priorities for communities, companies, employees, and government bodies. Georgia dropped from one to three in U. S. for business development due to “Quality of Life” issues. Remember crystalline silica is present in nearly all sedimentary and acidic rocks and basic igneous rocks are the perceived home of asbestos. These health issues, any other safety and health issues or perceived issues need to be resolved initially and any time they come up during life of deposit. When handled properly good safety and health programs add to the profitability of a deposit, not to the costs. Daily safety checks in operating plants and mines, and visual check of all drilling core with an eye for changes are a key to successful programs. Both plants I turned around had terrible safety records, too much airborne dust, and maintenance problems resulting in low productivity. In almost every case cost reduction was tied to eliminating maintenance and spillage problems which caused health problems and increased possibility of accidents. Implementation of these safety and health programs around 1970 starting prior to government programs, working with new MSA/MSHA regulations, working to improve the regulations as a president of a state mining organization, and maintain profits was a challenge, but very rewarding. Reducing time to make repairs and work simplification gleaned from observations and employee recommendations learned from making multiple daily inspections resulted in years without lost time accidents at two plants, reduced maintenance and 33-75 % improvement in cost per ton at A. P. Green Refractories Co. Eufaula Calcining Plant 1972-77 and Oil-Dri Ochlocknee plant and mine 1979-82. The Ochlocknee mine worked 1981-2007 without a LTA. The author participated in 16 acquisitions and new plant design projects. All of above reduced costs allowed mining deposits with higher overburden to be used and considered. Thereby increasing reserves considered in all operations."

Jan 1, 2016

"The 33rd annual American Mining Hall of Fame Awards Banquet and Fundraiser sponsored by the Mining Foundation of the Southwest (MFSW) will be held at the JW Marriott Tucson Starr Pass Resort & Spa on Saturday, Dec. 5, 2015. Each year, one living Hall of Fame inductee, two Medal of Merit recipients, several honorees from Mining’s Past, an Industry Partnership Award and a Special Citation are recognized for their respective contributions to the mining industry.Conrad E. Huss, cofounder of M3 Engineering & Technology Corp. located in Tucson, AZ, will be honored as the 2015 Hall of Fame Inductee and will serve as the keynote speaker at the awards banquet. Huss has served his company as chairman of the board for its entire 30 years and guided M3 to become a full discipline firm. Now M3 offers complete project execution through engineering, procurement and construction management procedure and has become a major engineer of large observatories around the world. The company now has more than 1,000 employees with offices in Arizona, North Carolina, Mexico, Argentina and is incorporated in Guatemala, Peru and Canada.The Medal of Merit will be awarded to Charles Michael (Mick) Lownds, founder of his eponymous consulting firm in Salt Lake City, UT, for his 42 years in commercial explosives following tenure at Sasol and at Orica. Originally from South Africa, Lownds immigrated to the United States in 1998 and has made significant advancement on electronic initiation systems. Lownds has published numerous technical papers, is an inventor with 12 awarded United States patents and was also the recipient of the President’s Award from the International Society of Explosives Engineers in 2000.Medal of Merit for young professionals will be awarded to Fiorella Giana, Freeport- McMoRan Inc.’s (FCX) mine manager at Bagdad. Born in Buenos Aires, Argentina, Giana left the small cattle farm where she was raised in 2001 to pursue higher education in the United States. For the past decade at FCX, as a result of her extraordinary combination of technical skill, leadership and commitment to excellence, she has been given positions of progressively greater responsibility from an entry level engineer to senior mine engineer, to chief engineer and now as mine manager, a position rarely held by woman."

Jan 1, 2015

By G. W. McCaa

Plans for a large industrial development in the Upper Ohio Valley between Moundsville and New Martinsville, West Virginia, were announced in the Fall of 1955. This development consisted of a large aluminum plant to be .built by Olin-Mathieson; a 675,000 KW power plant to be constructed by the American Electric Power Service; and a large coal mine adjacent to the power plant to be developed by the Hanna Coal Company Division of Consolidation Coal Company. The Ireland Mine was designed to efficiently supply fuel for the power plant, and this was the first consideration in the location of facilities, design, and development of plant and the selection of the equipment. The first problem was to find a suitable area for the surface facilities. A level bottom area south of the power plant was reserved for future industrial plants and this area was also blocked off from the coal reserves by an old abandoned and flooded mine. The coal storage facilities of the power plant were on the north side of the plant and it was decided to locate the coal mine preparation plant on-the shoulder of a hill above the power plant. This location required the excavation of 70,000 yards of rock but permitted the driving of a belt slope from the plant at an angle so that the haulage roads underground would miss the abandoned Cresap Mine. This location also permitted the plant to be located at an elevation so that the conveyors leaving the plant were horizontal and had sufficient clearance above the highway and railroad tracks. The other surface facilities consisting of air shafts, supply slope, supply yard, office, bathhouse, shop and warehouse were located in the valley some distance north of the preparation plant. This location placed these facilities one mile nearer to the working area. Plans for design and development of the mine were started in November 1955; grading and clearing of surface area was started in January 1956; the air shafts and slopes were started in the Spring and Summer of 1956; and the first coal was produced from the supply slope in October 1956. The preparation plant was put in operation in March 1958, and by July 1959 was producing 8,000 tons per day.

Jan 1, 1959

By Bruce Hebblewhite

"Historically there have been a number of different hypotheses and empirical models developed in an attempt to describe the nature of fracturing above longwall panels in underground coal mining. The motivation for such models varies, ranging from understanding the impact of mining on surface subsidence to back-analysis of caving behaviour in the immediate roof behind the longwall face. One of the most critical motivating factors that is taking on increased importance in many coalfields is the need for better understanding, and hence prediction of the impact of mining on overlying strata, particularly strata units acting as aquifers for different groundwater horizons.This paper reviews some of the major prediction models in the context of observed behaviour of strata displacement and fracturing above longwall panels in the Southern Coalfields of New South Wales, south of Sydney. The paper discusses the parameter often referred to as “height of fracturing” in terms of the critical parameters that influence it, and the relevance and appropriateness of this terminology in the context of overlying sub-surface subsidence and groundwater impact. The paper proposes an alternative terminology for this parameter that better reflects what it is and how it is used. The paper also addresses the potential role of major bedding shear planes mobilised by mining and their potential influence on overlying subsidence and groundwater interference.INTRODUCTIONKnowledge of the detailed nature of rock deformation and failure above any form of large-scale underground mining is always going to be limited to interpretation from a very incomplete set of data. It is extremely difficult, if not impossible, to directly measure the detailed nature of the rock failure, fracture networks, and deformational behaviour above an extracted mining area. Limited techniques such as borehole extensometry can provide some evidence of relative or incremental deformation in the direction of the borehole (usually vertical). However, such data cannot assist below the horizon where full caving has caused major rotation and dislocation of rock blocks and effectively destroyed the instrumentation borehole. Above such a horizon, the data is only valid along the axis and in the direction of the borehole, and to the level of detail defined by the extensometer anchor spacing intervals. It is often also restricted by major shear planes intersecting the extensometer borehole and cutting off the instrumentation below the level of the shear horizons."

Jan 1, 2019

By Kiseong Hyeong

Geochemical characteristics, shapes and surface textures of manganese nodules from two blocks located in the southern (9°N~11°N) (KR5) and northern parts (16°N~17°N) (KR2) of the Clarion-Clipperton fracture zone between 130°W and 133°W were investigated in terms of sediment distribution and metal contents for better understanding of controlling factors of seabed manganese nodule genesis. Two blocks, ~700 km away, are dominated by the nodules with different origins, each of which has distinctive chemical compositions and shapes. KR2 block is dominated by irregular and poly-type nodules of hydrogenetic origin with low Mn/Fe, Cu, and Ni contents and high Fe and Co contents, while KR5 block is dominated by discoidal and ellipsoidal nodules of diagenetic origin with high Mn/Fe, Cu, and Ni contents and low Fe and Co contents. In addition, nodules from two blocks show different surface textures: smooth and transitional in KR2 and dimorphic and rough in KR5. Sediment distribution is also distinctly different in two blocks. The study area consists of three sediment types: Unit 1 ? dark brown mud with an intermediate metal content, Unit 2 - yellowish brown mud with poor metal content, and Unit 3 ? black metalliferous mud. KR2 is characterized by the thick occurrences (<250 cm) of Unit 1 at the surface while Unit 1 distributes thin (<12 cm) in KR5. Higher primary production rate in the southern block proximal to equator and resulting high supply rate of metals to the sediment might be the cause for the dominance of diagenetic nodules in the southern block, but potential controls of different sediment distribution on the nodule genesis can not be excluded. If it is the case, Unit 1 with an intermediate metal content likely has been acting as a sink for metals by precipitating micro-nodules within the sediment column. Within the KR5, the hydrogenetic nodules occur preferentially on the seabed where Unit 3 distributes at shallower depth (<50 cm). Unit 3 has 2 or 3 times higher metal contents than other types of sediments. It might play the same role as Unit 1 in the northern block that prevents the supply of metals to the sediment surface. It also supports the potential effect of sediment distribution on nodule genesis. Keywords: Nodules genesis, Sediment distribution, Supply rate of metals

Jan 1, 2008

By M. K. McCarter

INTRODUCTION Mining activity and basic geometry of open pit mines represent major obstacles in geotechnical mapping. Mining creates a constantly changing topographic base which may require corresponding updates of field sheets before mapping can proceed. The rapid pace of mining operations may also threaten removal of important geologic and engineering features before adequate base maps and surveyed control can be established. Maintaining such control during the mapping process is yet another complication. In addition, basic geometry yields nearly vertical outcrops. Access required for direct observation and measurement, essential to conventional plane table or Brunton techniques, may be impossible on such surfaces. Also, nearly vertical surfaces appear fore- shortened in plan view. This reduction in area provides little room for recording geotechnical detail. The difficulties encountered in open pit mapping can be substantially reduced by utilizing a modified form of plane table photogrammetry. This technique employs photographs taken from the ground as the base upon which details are recorded. The information is later transferred to plan maps using the plane table principal of location by intersection (Davis and Foote, 1953). Any change in topography can be accommodated by simply taking additional photographs. Proper spatial representation in plan view can be determined later when updated maps and surveyed control are available. In addition, horizontal photographs from the ground portray vertical surfaces as they are normally viewed thus providing maximum area for recording details. Any detail which can be identified on the photograph can be located correctly even though the actual feature is physically inaccessible. The technique presented in this paper was developed by the author for special applications at Kennecott&apos;s Bingham operation in 1970. The procedure was rapidly adopted for general use and has proven to be a practical tool for detailed investigation of potentially unstable areas (Zavodni and McCarter, 1976) and routine geological mapping (Richardson, 1975). Major advantages, other than those mentioned above, include: The speed and accuracy are greater than for conventional plane table or Brunton methods. The photo- graph represents a view from a station at some distance from the area being mapped. This vantage point allows evaluation of possible continuity of structures from one level to another which may not be evident at close range. The photographic base provides a permanent record of the topography at the time

Jan 1, 1977

By Rush H. Spedden

Antoine M. Gaudin was a vigorously creative man and throughout his career an internationally respected leader of his chosen profession of mineral engineering. To his professional colleagues and clients and to his former students he was both "Professor Gaudin," their wise, inspiring, and enthusiastic teacher, and "Tony," their warm and beloved personal friend. Thus, his teachings and his unique inquisitive-creative approach to problems live on and grow in the accomplishments of his students. Antoine Gaudin was born at the turn of the century in Smyrna, Turkey, where his father was general manager of a French-owned railroad. It was there that he spent his earliest years, literally in the shadow of classic marbles from diggings at the site of Aphrodesias, which his father had rediscovered. From this, Tony developed an appreciation of the arts. He eventually surrounded himself with art treasures; he reveled in the beauties of nature, and he cherished his choice collection of rhododendron and azaleas. His artistic traits led him to take up painting when his daughter, Elinor, sought his help in blending her tints for a water color. In typical fashion he was not satisfied until he had become adept with pastels in capturing the muted tones of distant landscapes and the technical properties of light and color. Unfortunately not many paintings were produced. His paternal grandfather was Marc-Antoine Gaudin, for many years secretary of the Academie des Sciences in Paris, and a well known chemist and mineralogist whose primary interests included the synthesis of precious stones. In an autobiographical sketch written in 1939, Dr. Gaudin recalled: "As a child I saw many of the original synthetic sapphires and rubies which my father had inherited. This, coupled with my father&apos;s avocation of mining ventures and archaeology, early decided that I would become connected with the mineral industry." One can speculate that the archeological activities of his father, the mineral systhesis experiments of his grandfather, and his own early studies of mineralogy and geology formed a base from which his special talents as an experimentalist were to develop. Laboratory research under Professor Gaudin&apos;s personal guidance was often an intensive, fast- moving, and exciting experience--especially to a new graduate student. First of all, Professor Gaudin was a highly skilled observer with seem-

Jan 1, 1976

By Frank A. Aplan, Barbara J. Arnold

INTRODUCTION Coal flotation is a fine cleaning process usually present in all but the older preparation plants. It is typically used to clean either the -28 mesh (-500 pm) or - 100 mesh (- 150 pm) raw coal fines, and it is the only commonly used preparation method that can effectively recover coal particles finer than, say, -200 mesh (74 pm). To place coal flotation into its proper context, it should be noted that the bulk of the raw coal sent to preparation plants is treated by gravity concentration. On a worldwide basis, from 0 to 40% of the raw feed to a preparation plant is treated by flotation, while in the US only about 10% of the raw coal feed is floated. Roughly, 50% of US preparation plants use flotation to process at least a portion of the raw coal feed, although most of the newer plants incorporate flotation into their flowsheet. Historically, the process has only been used to treat the raw coal fines naturally occurring after mining and a minimal amount of crushing at the preparation plant. This is in sharp contrast to ore flotation where, typically, the entire tonnage is first ground and then floated. As a consequence, of the approximately 500 Mt of raw coal treated by preparation in the US yearly, it is estimated that only about 40 Mt are sent to the flotation circuits (producing -30 Mt of clean product).&apos;O1 This small tonnage compares with about 400 Mt of ores treated annually in the US by flotation. By analogy to ore flotation, where it is an effective, large tonnage, low cost, easy to operate process, its potential in coal preparation should be great. Reasons why it has not achieved a greater potential in coal preparation have been attributed by Aplan 102 to: 1. Indifferent results in removing sulfur and ash. 2. Difficulty in floating low rank and oxidized coals. 3. Inability to make clean separations with the finer sizes. 4.Inability to clean slurries with a high percentage of clay. 5. Problems with water chemistry. 6. Competition with effective, gravity coarse-cleaning methods. 7. Low value of coal historically. 8. Dewatering problems. 9. Transportation and storage problems with coal fines. 10. Marketing problems with coal fines. 11. Inadequate attention to the process in the plant. 12. Inefficient use of capital and operating monies. Inadequate attention to procedures, machines, and processes.

Jan 1, 1991

By Steve Kral

The mining industry has waited a long time for the resurgence in commodity prices and demand for its products that it is now experiencing. Base and precious metals prices have reached levels not seen before. Coal prices have also shot up. And there are encouraging signs in the uranium market. China’s nearly out-of-control economic growth is said to be the leading driver for much of this resurgence. But demand for energy, metals and building materials in the United States and elsewhere are also major factors. The result is that expansions and new mines for nearly every commodity are in the works.

Jan 1, 2005

By L. L. M. Kadnuck

The U. S. Bureau of Mines studied the suitability of using flexible tendon cable supports as secondary roof supports at a western coal mine. The supports were installed in longwall tailgate entries and grouted into the mine roof. This paper quantifies how the use of cable supports can improve ventilation in underground airways. A mine ventilation simulator was used to model the changes in airway resistance and airflow and show potential for further ventilation improvements when cable supports are used. Model results indicated that cable support use translated to savings in air pressure around the longwall panel and increased airflow at the longwall face.

Jan 1, 1994