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By Sanna Mustonen, Seppo Mellanen, Mellanen, Pekka Kantia

In Finland Posiva Oy (Company) is responsible for implementing the final disposal programme for spent nuclear fuel of its owners. Company is preparing for the geological disposal in Olkiluoto crystalline bedrock. Excavation will be executed using drill and blast method. Method unavoidably causes variable excavation damaged zone (EDZ) on tunnel surfaces. The EDZ constitutes a risk for long term safety and therefore good control methodology of the blasting technique is required. The thicker the EDZ, the greater is the likelihood for detrimental transport pathways for ground water to transport radio nuclides, should those be present.

Jan 1, 2016

By Andrew Cormier

"The Early YearsThe Grand Lake coal fields in New Brunswick undoubtedly can claim the distinction of being the first part of North America from which coal was mined and exported. The diaries of the Governor of Massachusetts during the first half of the 17th century clearly establish this fact. There is also a reference in the Massachusetts Colonial Records which indicate that coal exports to New England were carried on as early as 1639.It has generally been believed that Cape Breton mined and exported the first coal; but no account has yet been found to show that this took place earlier than 1676. Both at Grand Lake and Cape Breton the coal was near the surface and removed by hand.When the country passed entirely into British control, Joseph Garrison, a grantee of Maugerville in 1765 moved to Jemseg. While surveying the north side of Grand Lake he discovered outcroppings of coal on the shore of the lake and further inland along the banks of streams. It was he who named Newcastle Creek in reference to Newcastle in England, which was widely known for its coal mines. Joseph Garrison died in 1783, the same year that an influx of United Empire Loyalists increased settlement in Newcastle Creek in the area now known as Minto. Many Loyalists also settled on the shores on both sides of Grand Lake.At a number of points where streams cut channels through the overlying strata, the coal seams were exposed and it was at such outcroppings that the early operations were carried on. A tunnel was driven in the seam from the foot of the bank, affording both an outlet for the coal and natural drainage from the workings. The coal was loaded into boxes and wheeled over wooden tracks to a point on the surface where it could be dumped into wagons and hauled either direct to destination or to where it could be loaded into boats and conveyed via Grand Lake, Jemseg and Saint John river to points of distribution.In the latter part of the 18th century and first years of the 19th century, farmers were undisturbed by mining the coal found on their lands. This was soon to change and resulted in many years of contention between landowners and the Legislative Assembly.New Brunswick became a separate province from Nova Scotia in 1784 and the seat of government moved from Saint John to Fredericton in 1788. The new province had its first priorities; the establishment of counties, the building of roads, and the development of its natural resources. From 1805, records of coal and land development are fairly complete. These records show the increasing interest of the Province in developing the coal fields. This interest resulted in the granting of coal leases to many friends of the government, without any regard for the wishes of the farmers who owned and had occupied the land for many years. The farmers fought back with lawsuits and other forms of dissension but with little success. By the end of the 19th century there were very few farm operators left in the area."

Jan 1, 1992

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

Hydrodynamic investigations in exploratory boreholes and grouting holes are conducted for the purpose of obtain¬ing information about the hydraulic properties of the hydrostratigraphic section to be intersected by the proposed underground workings. The information obtained from the investigations provides the basis for calculating the hydrau¬lic coefficients of fractured permeable rock, the dimensions of the anticipated grout isolation curtain(s) around the un¬derground workings, the number and location of grouting holes, the injection pressure modes, and also the volume(s) of grout that will be required (Anon., 1976, 1978). The following data on each aquifer are obtained from the investigations conducted in monitoring and grouting bore¬holes and the analysis of the results: 1) the top of each hydrostratigraphic unit, 2) the thickness of each unit, 3) the ground water fluid potential distribution in each unit, 4) the coefficient of permeability, 5) the piezoconductivity, 6) the fracture porosity, 7) the geometry of the fractures in the rock, 8) the elasticity-compressibility coefficient of the fractured rock, 9) the chemical composition of the ground water, 10) the direction of flow of the ground water, and 11) the expected inflow rate of water into the shaft, drift or tunnel. STG uses its DAU-3M type flowmeter to conduct in¬vestigations of directions of flow in vertical, inclined and horizontal drillholes. The DAU-6 instrument is used to de¬termine the direction of flow of ground water in each frac¬ture or fractured aquifer. Various singular and double DAU type packers are used for pumping and for injection studies (tests) and for flowmeter investigations. Normally the instruments enumerated above permit in¬vestigations to be conducted in each separate aquifer with¬out reinforcing the holes with casings. On the basis of these investigative data, both the hydraulic properties of unfractured rock and the hydraulic properties of the fractured rock are estimated. Dual porosity rocks require special attention because they tend to segregate the grout. 3.1 FLOWMETER INVESTIGATIONS IN BOREHOLES The STG flowmetric methodology is based on the mea¬surement of the ground water flow rate through the borehole by hydrostratigraphic interval after the disturbance of the hydrostatic equilibrium in the "hole-aquifer system" (after pumping or injecting). The relationship of the head changes to the discharge into or from a particular hydrostratigraphic unit obtained during the tests serve as the basis for calcu¬lating the hydraulic properties. Flowmetric investigations facilitate the determination of the number of aquifers, their depths, their thickness, the hydraulic properties of the fractured rock and the magnitude and direction of the flow of ground water. 3.1.1 FLOWMETER HARDWARE STG conducts flowmetric investigations in boreholes using its DAU-3M-108, DAU-3M-73, DAU-3M-57 and DAU-3M-44 instruments.' They have respective external diameters of 108, 73, 57 and 44 mm. The type of flowmeter selected for use depends on the borehole geometry and the technological scheme for carrying out the investigations. Boreholes with a drilling diameter of 76-93 mm are inves¬tigated with the DAU-3M-73 flowmeter; boreholes drilled by bits with a diameter of 112 mm and more are investi¬gated using the DAU-3M-108 flowmeter. The DAU-3M¬108 and DAU-3M-57 instruments are used for flowmetric investigations with a packer. 3.1.1.1 The Downhole Sensor The sensor design of the DAU-3M-73 hole flowmeter is shown in Fig. 2. The design of the DAU-3M-108 instru¬ment is similar to the design of the DAU-3M-73 instrument. The frame of the flowmeter sensor shown in Fig. 2 consists of a casing, an upper and lower centering mount and two rings to which the guiding rods are attached. The upper rods are built into the connector bushing; the lower rods are built into the coupling sleeve. The borehole cable is attached using a half-coupling, a packing ring and a constriction nut. Thus, the frame of the flowmeter sensor is made so that the free passage of water to the impeller is facilitated along with the necessary rigidity. The primary moving component of the flowmeter is the double-bladed impeller, which rotates on cobalt-tungsten pivots and agate thrust bearings. Special extended air cham¬bers protect the supports of the impeller from the action of the borehole fluid which may contain fibrous and abrasive particles. The air located in the chambers shields the sup¬ports from direct contact with the borehole fluid when the sensor operates in a borehole. The hollow casing of the impeller serves the function of a lower cap. The upper cap is attached to the casing using a threaded connector; it is affixed also with a lock-nut. An adjusting screw with a

Jan 1, 1993

By Jay Keng

During the construction of shaft on the fractured and weak rock strata, the stability of construction plays an important role. This study includes earth stress estimation, support method, and stability analysis. Stresses on the site of three shafts include structural residual stresses, earth pressures, and hydrostatic pressures were determined. The basic method of estimating the support strength is dependent upon the' size of the plastic ring after rock excavation. During the construction of the shaft, geotechnical measurements, physical properties of the rock, and the change after excavation have to be evaluated again. Geology and construction methods need further consideration.

Jan 1, 1988

By J. E. Tobey

UNDER war conditions coal immediately assumes a position of highest importance for coal must carry the basic load for industry. The upward trend in production continued through 1941. Bituminous coal production alone about equaled the total output of bituminous and anthracite in 1940. The figures for 1941 approximated 500,000,000 tons for bituminous, and 54,000,000 tons for anthracite, and production of by-product and beehive coke exceeded that of the previous year. The coal industry, under the National Bituminous Coal Act, is going through a period of readjustment, and is beginning to stabilize itself under the statute. The Federal Mine Inspection Nill became law during the year.

Jan 1, 1942

By D. A. Trotter

"This paper traces the main advances in mine lighting from neolithic times to the 1960s. The earliest lights consisted of reeds or strips of wood soaked in grease. As man's knowledge increased, the oil lamp, the candle, kerosene, acetylene and finally electricity have allowed him to see where no natural light could penetrate. These advances in technology, however, did not come easily. Many false starts were made and great dedication was required by a few lighting pioneers to overcome the many technical difficulties that were encountered. The lighting of coal mines in particular caused unique problems, as there was the ever-present threat of methane explosions. Advances in mine lighting will continue, but these advances are only possible because men like Davy, Wolf, Marsuat and Edison paved the way.IntroductionMan's first efforts at mining consisted of merely picking rock off the surface of the earth. The sun was his only source of light. The early mines of the Neolithic age were simply vertical pits, 5 to 8 m deep. In order to follow an orebody, miners began to dig horizontally from the bottom of these pits, and as work progressed the need for additional light became apparent. With the discovery of the light-giving properties of fire, man was no longer dependent on natural light. Figure 1 gives an impression of the earliest artificial mine lighting used.Early Evidence of Mine LightingCarbonized residue from wood chips which had been inserted into rock cracks can be found in early Neolithic workings. The mines at Krzemionki Opatowski in Poland contain very narrow low tunnels that were lit in this fashion and were worked from 2300 to 1800 B. C. The wood had to burn downward to provide light and because wood burns poorly in this position, the chips were probably soaked in animal fat to make them burn more easily."

Jan 1, 1982

By G. Townsend Harley

STOPING methods in which shoveling plays an important part are gradually being replaced by other and cheaper methods. But there will always be considerable shoveling done underground in stopes as well as in drifts, tunnels, urinzes, and shafts. At the mines of the PhelpsDodge Corporation at Tyrone, N. M., the cost of shoveling in all stopes in 1917 amounted to 24c. per T. In the top-slice stopes for the same period, it cost 27 c. per T. or 16 per cent, of the total cost of these stopes. The tonnage for shovelers from all stoping was 9.3 T. per man, and for topslicing 8.2 T. per man. These stopes were not unduly hot, and there was not more than the usual amount of timber to interfere with the work of the men. The tonnages obtained per shoveler were considered low; first, because of a poor grade of Mexican labor, many of the men having come in from railroad grading camps; and second, because of a poor spacing of raises, especially in the top-slice stopes, where, in general, they were spaced 25 ft. by 66 ft. (7.6 by 20.1 m.) centers. The average wage per laborer shift was $2.67 during the year. It was thought, however, that even under these conditions the men were not producing the tonnage that they should, so, with the consent of the management, the writer undertook to determine how the general efficiency of the underground shoveling could be improved. No predetermined plan for conducting these experiments was arranged because we had no definite ideas as to the scope of the work or the number of elements into which the investigation would resolve itself, before all of its phases could be determined. We were sure, however, that any work that would thoroughly cover the ground would have to be in the nature of a systematic time study, combined with a course of instruction in correct shoveling methods and adequate and intelligent supervision of the work. Two or three companies in the Southwest have done some work to determine the proper shovel to be used undcrground, but so far as is known the work has been limited to equipping certain parts of their mines with a particular type and size of shovel and thereafter watching the cost and efficiency records. In each case it seems to have been the shovel that held the 21-lb. (9.5-kg.) load that gave the bcst results. Excepting personal communications from these companies, the only data available on scientific shoveling are contained in F. W. Taylor's book, "The Principles of Scientific Management," and D. J. Haucr's article, in The Contractor for March 29, 1918, "A Hundred Hints for Shovelers." This paper discusses and draws conclusions from several thousand time-study readings, taken both on the surface and in the mines for nearly a year. A sufficient number of readings were taken on each factor in the problem for the plotting of curves of the performances and to obtain accurate indices of the work to be expected from this class of labor. The work is not as complete in all of its details as we would like to havc it, because we were forced to stop the work temporarily, owing to the numbcr of men going into the National service and our inability to get others who could make time studies. The results obtained so far, however, have been of such a startling nature that we have decided to submit them at this time, subject to future modification. It is hoped, also, that a free discussion of this paper will lead to a disclosure of any errors that may have becn embodied and offer some valuable suggestions for the conduct of future work.

Jan 1, 1920

By R. B. Hoy, Stanley J. LeFond, Unni H. Rowell, K. Reckling, Derek G. Fullerton

In 1989 natural industrial diamonds counted for 55% of the world's natural diamond production. Australia is currently the leading producer (35%). Zaire is the second largest producer (19%). of what is primarily industrial grade rather than gem grade. Botswana (17%) is third, with the former USSR (15%) fourth, and the Republic of South Africa (8%) fifth. Diamonds are also mined in Angola, Namibia, the Ivory Coast, the Central African Republic, Ghana, Tanzania, Guinea, and other African countries. In the Western Hemisphere, Brazil is the principal producer, with some production from Venezuela and Guyana [(Fig. 1)]. A very small output of diamonds is mined today in India, which was the first source of commercial production. In the United States, efforts at commercial diamond mining have been confined to a small area near Murfreesboro, AR. The first diamond was found in a pipe there in 1906. Small scale trial mining has not, however, proved economical. Since diamonds were first discovered more than 2,000 years ago, only about 380 t have been mined. In order to obtain 1 g (5 metric carats) of diamonds, it is necessary to remove and process approximately 25 t of rock. Recovering this small percentage involves a combination of highly developed techniques in mining and extremely sophisticated processes in diamond recovery. END USES Diamonds are used for two unrelated end uses: gem diamonds are jewels of great beauty, while industrial diamonds are essential materials of modem industry. Although imitation stones are substituted for the gem diamond, none of these matches its properties sufficiently well to offer real competition. Synthetic industrial diamonds are now of a quality and size that permit them to be substituted for natural diamonds in numerous industrial applications. For example, synthetic diamonds are available today in sizes up to 100 stones per carat (1.2 to 1.4 mm). In addition, polycrystalline synthetic diamond inserts, such as De Beers Syndite", General Electric's Compaxa and Stratapax", and Megadiamond's Megapax" have replaced natural diamonds in turning tools, mining and oil drilling bits, and dressing tool applications. Industrial Diamonds The diamond is by far the most important industrial abrasive. As recently as 50 years ago, consumption of industrial diamonds was less than that of gem diamonds, but since that time, industrial use has grown to a position of great dominance. During the six-year period 1929 to 1934, the material produced for industrial use amounted to about 74% by weight of the world's total output of diamonds. In 1989 the percentage of natural industrial diamonds mined in the world was 55%. When synthetic industrial diamonds are added to the natural industrial diamond figures, this percentage becomes 87% of total world diamond production including gems, near gems, industrial, and synthetic stones. The many uses responsible for these significant increases are dependent on the properties of the diamond, including hardness, cleavage, and parting, optical characteristics, presence of sharp points and edges, and capacity for taking and maintaining a high polish. The utilitarian role of the diamond was confined primarily to lapidary products until the industrial revolution, which created the first demand for diamond as an industrial tool. In 1777, a British inventor and instrument maker, Jesse Ramsden, used a diamond to cut a precision screw for an engine that he had invented. The first authentic description of industrial diamonds being set in a saw was recorded in 1854 by a Frenchman, Durnain. Eight years later a Swiss watchmaker, Jean Leschot, developed the first diamond drill bit for use in a hand operated machine, which was employed to drill blastholes in rock. In 1864, diamond bits were put to their severest test up to that time in the construction of the Mont Cenis Tunnel in the Alps. A few years later a steam-powered diamond drill with a speed of 30 rpm was able to penetrate rock at the modest rate of 0.3 m/hr. As the industrial revolution gained momentum on both sides of the Atlantic, metal replaced wood and machines replaced people. Thus the foundation was laid for precision engineering and the recognition of diamonds as an indispensable tool of industry. The next major demand for industrial diamonds came after World War I with the development of cemented carbide cutting tools. Diamond was found to be the most effective medium for finishing and grinding the new ultrahard metal. This discovery rapidly increased the demand for industrial diamonds. The availability of inexpensive diamond material inspired tremendous research into applications. By 1935, the first successful phenol-resin grinding wheel containing diamond had been marketed. Soon afterward, the metal-bonded and vitrified diamond wheels were produced, and, as the matrices and bonds that held the diamond grit in place began to improve, the popularity of diamond grinding wheels grew. The outbreak of World War II, and the accompanying increase in use of hard-metal tools in the munitions industry, increased the demand for industrial diamonds. Since about 1950, the development of ultrahard ceramics, semi- conductor materials, plastics, and exotic metal alloys has further consolidated the diamond's position as an indispensable tool of industry. Only diamond is hard enough to cut these superhard materials with the precision, speed, and economy that industry demands today. Special machines equipped with industrial diamonds are used to remove bumps from concrete runways and highways and to modify highway surfaces in order to prevent skid accidents. Many skids are caused by hydroplaning, a phenomenon that occurs when the roadway is wet. Tires mount a film of water and virtually lose contact with the road surface. Diamond machines cut neat, narrow

Jan 1, 1994

By C. Kwong, L. H. Rorden, T. C. Moore

This report describes the feasibility test of an electromagnetic method for locating, boreholes, with or without casing, that have been drilled in conjunction with mining operations but whose location with respect to the tunnel is not precisely known. The first tests were conducted with the cooperation of the Kerr-McGee Corporation, Church Rock Mining Operations, New Mexico, in September 1977. These tests demonstrated that accuracies of better than ±15% of range and ±1° of bearing can be achieved with a simple system and setups. Subsequent tests, in January 1978, on a horizontally drilled pipe line river crossing application, by Ti tan Contractors of Sacramento, California, demonstrated that accuracies of better than one-half percent of range and a few tenths of one degree of bearing can be achieved by using refined techniques. The demonstration tests were done at ranges on the order of 100 to 200 feet but the method is capable of working at much greater distances.

Jan 5, 1978

By E. B. Olds

This paper describes the methods and costs of deepening the 3-compartment underground shaft in the Hooper Tunnel adit at the Bunker Hill & Sullivan Mining and Concentrating Company Crescent mine in the Coeur d'Alene mining region, 7 miles east of Kellogg, Shoshone County, Idaho. The work was done with the aid of a Defense Minerals Exploration Administration contract (No. Idm-E493, Docket No. DMEA-2764). Preparation for deepening the shaft was begun March 26, 1953, and sinking was completed May 27, 1954. The shaft, bottomed at 1,275.5 feet in 1939, was deepened to 3,276 feet, using mechanical mucking and intermediate pockets. A total of 50,349 man-hours was required to complete the work of sinking 2,000.5 feet, cutting 7 pump stations, and driving 2 pocket raises. The rate of advance was 7.1 feet of shaft per operating day, or 8.4 feet per day of actual sinking. The total direct cost for sinking was $205.20 per foot.

Jan 1, 1957

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION The two unit operations in cut-and-fill stoping that would benefit most from increased mechanization are drilling blastholes and moving broken ore to the ore¬passes. Drilling in current practice is done with hand¬held jackleg and stoper drills. Broken ore is commonly moved from the face to the orepass by a small slusher. This type of equipment can be easily moved through the raises. Hand-held drills have a slow drilling rate and require one operator for each machine. Drilling can be done much faster by small jumbo-mounted machines as two or more drills can be operated by one miner. Load-haul-dump (LHD) units move broken ore from the face to one or more orepasses at a much faster rate with less loss and dilution than slushers. These units are mobile and may be used at more than one place during a shift. The problem which must be overcome in applying this higher degree of mechanization is the restricted means of access inherent in current practices. Cut-and¬fill methods are usually applied to mining veins that are narrow, irregular in dip and strike, or where ground con¬ditions require the support of sand fill. The mechanized equipment which is needed to accomplish high produc¬tion cannot be raised or lowered through raises without being disassembled. A jumbo which will give the desired performance would be mounted on a rubber-tired carriage with a diesel engine for propulsion. Twin hydraulic booms run by a small air motor would carry pneumatic drills. Over¬all dimensions of the unit would be about 8 m (26 ft) in length, 1.6 m (5'/z ft) in width, and 1.6 m (5'/z ft) in height. An LHD of the desired capacity would be rubber¬tired and either diesel or electric-powered. The overall dimensions of a 2.3-m (3-cu yd) unit will be about 8 m (26 ft) in length, 1.8 m (6 ft) in width, and 1.5 m (5 ft) in height. It will weigh about 12 700 kg (28,000 1b). It can readily be seen that conventional raises can¬not handle this size equipment. The most desirable means of access enables equipment to be driven into the stoping areas. Ramps can enter the stope by being driven upward in the vein walls or by being developed in the vein on top of the fill as the stope advances. This ramp system not only permits the use of rubber-tired equipment in multiple working places but also enables its removal for repairs without component breakdown. In the proposed method, as in conventional cut-and¬fill stoping, sand fill will replace the mined-out ore. By constructing an inclined timber floor at one end of the stope, a ramp can be maintained on top of the fill and advanced as the stope is advanced cut by cut (Fig. 1). The ramp is, therefore, developed during the stoping process and is only driven as a tunnel when rising over and around the sill and crown pillars. A stoping section consists of three blocks of different geometric shapes. The first and last blocks to be mined will be triangular in the plane of the vein, and the blocks between will have the shape of parallelograms (Fig. 1). The first block will develop the initial ramp by breast stoping and filling to the ramp floor (Fig. 2). The ramp will need to be completed through to the next level be¬fore stoping of the subsequent block can begin. The subsequent blocks are mined by back stoping with the next ramp being developed at the other end of that block (Fig. 3). The length of each subsequent block can be varied without changing the slope of the ramp. As the new ramp is developed, the previous ramp is filled along with the stope cut. Access up and down is maintained through the stope. The final block is again breast stoped as no new ramp will be developed at the end line for jumbo egress. Boreholes provide ventilation between levels as well as utility access. They can also provide a secondary escapeway for personnel during periods of development. These boreholes would be located about 3 to 4.5 m (10 to 15 ft) in the footwall and spaced about 183 m (600 ft) apart along the vein. The orepasses are spaced about 61 m (200 ft) apart in the vein. This spacing allows not only a short tram for the LHD unit but also provides surge capacity for the ore between haulage trains. These passes are raised from the haulage levels. As development progresses, the raises are driven through the sill pillar, and then car¬ried in the fill as stoping progresses. Ventilation and utilities can be maintained through an orepass when it is not needed for ore. The end line orepasses are intended for this purpose as well as for personnel emergency egress. Overall dimensions and criteria assumed to illustrate the ramp method of access in a highly mechanized cut¬and-fill stope are: haulage level spacing of 61 m (200 ft); ramp gradient of 17%; average dip of approximately 1.57 rad (90°); initial stope length of 358 m (1177 ft); subsequent stope length of 179 m (588 ft); average stoping width of 3 m (10 ft) ; sill pillar thickness of 3.6 m (12 ft); crown pillar thickness of 1.8 m (6 ft); bored raises of 1.5-m (5-ft) diam; and cribbed orepasses of 1.2 x 1.2 m (4 x 4 ft) through the sill pillar and 1.2 m (4 ft) square inside the cribbed portion. DEVELOPMENT The ramp access system lends itself well to conduct¬ing the several phases of development concurrently. Haulage Drifts and Raises The first development step will be to drive the 3 x 3-m (10 x 10-ft) haulage drifts above and below

Jan 1, 1982

By Louis A. Panek

INTRODUCTION Block caving is distinguished from other types of mining in its requirements that the mineral body to be extracted must cave after a relatively large tabular open¬ing is created underneath it, and that the caving must be predictable and controllable for both operating and safety reasons. A mineral body that caves readily may present special problems with respect to ground support in the mining access openings, where stability, rather than failure, is the desired objective. The complete ex¬traction of a caving block, which may measure 50 to 100 m on a side and 100 to 300 m in height, tends to create a large void, resulting in significant changes in the ground stress system and a migration of rock from all directions toward the void; these events begin very early in the life of the mine and lead to additional spe¬cial problems with respect to control of ground move¬ments in the mine working areas, in the rock adjacent to the caving stope, and on the subsiding surface above the mine. Thus, the mechanics of an undercut-cave operation involve a broad range of geotechnical subject matter, including the determination of the lithostatic earth stresses, the structural properties of a large jointed rock mass, the flow properties of the fragments after caving, the stability of tunnel-like openings, and the stability and displacements of the rock structure (1 to 2 km in size) that surrounds the ore body. The objectives are achieved by in-place measurement of rock properties, laboratory tests, theoretical structural analyses of the related prob¬lems, measurements of structural behavior in and above the mine, and comparison of theoretical predictions with actual performance in order to interpret and refine the predictive capability. Undercut-cave mining (Julin, et al., 1973) differs from most other methods in that the caving is an essen¬tial and inherent part of the excavation process and in¬volves the movements of millions of cubic meters of rock over distances of hundreds of meters, as compared to the gradual 2-m settling of the overburden typical of full¬caving longwall coal mining. Undercut-cave mining may involve a rock mass on the scale of a 1-km diam open¬pit mine, but mining is several times deeper than that in the open-pit mine; furthermore, a large portion of the rock mass is in a failed, rather than a stable, state. Ground-control problems are inevitable in a situa¬tion where ground movements are occurring on such a large scale, and where the rock mass is expected con¬veniently to cave at one location but to be stable at an¬other. To devise effective responses to these ground¬control problems, one needs to know first of all what is happening, not merely at the point where the symptoms are observed, but more broadly with respect to the be¬havior of the entire mine rock-mass structure over the total area of extraction, including the effects on the sur¬face as well as those underground. Monitoring structural behavior on this scale would require a sizable group of people, if they attempted to measure all the significant mining-induced changes of ground displacement and pressure, even on a 100-m grid of points. The practical approach is to conduct the mea¬surements activity by a judicious sampling, making in¬termittent, concentrated efforts in selected areas of lim¬ited extent to study problems of immediate importance; simultaneously a minimum level of surveillance is main¬tained over the total structure to delineate long-term be¬havior trends in the major structure, which provide the baseline reference data that are so essential for interpre¬tive purposes. Mine-structure behavior depends not only on the mine extraction configuration and sequence, but on the rock structural properties and how they change from one location to another in the mine. Increasing attention is being given to determinations of rock strength and de¬formation by performing mechanical properties tests in place and in the laboratory, and by drawing inferences from the mapping of geologic structure in mine openings as they are driven and from the logging of cores ob¬tained from exploration drilling that may be performed primarily for delineating ore reserves. Monitoring of mine-structure behavior and inter¬preting it in terms of the growing excavations and the structural properties of the rocks have significant im¬pacts on designing the major undercut-cave operations of undercutting and drawing the ore, supporting the ac¬cess openings, and controlling surface subsidence. The purpose of this chapter is to call attention to the com¬mon ground-control problems encountered in undercut¬cave mining, to relate them to the mining process, and to explain the geomechanical approaches to analyzing these problems with emphasis on methods that are most readily available to the mine staff. UNDERCUTTING, CAVING, AND DRAWING A distinguishing characteristic of an undercut-caving method of extraction is the objective that the rock mass fails of its own weight in a predictable and controllable manner, after it is undercut by excavating a tabular¬shaped void, horizontal or inclined. Induced or forced caving refers to procedures that may be resorted to when the ore will not self-cave over the desired span. Block caving, panel caving, and mass caving are terms used to denote the particular sequence of undercutting and ex¬tracting the total area of the ore body on a single min¬ing level. The discussion herein is not specialized to any of these variations; the word block as used does not imply anything as to sequence. CREATING THE UNDERCUT A common undercutting method is to drive a series of drifts and blast the intervening pillars. Theoretical structural analysis and practical experience show that ground-support problems in the undercutting drifts and in underlying access openings as well (the same princi¬ple applies even if undercutting is done by belling out drawpoints from the slusher drifts) are minimized by retreat-mining across the undercut slot according to a sequence such that the growing undercut area retains a

Jan 1, 1982

Improvements in cost efficiency, noise control, environmental design and ergonomics ? the mining and construction industries are facing demands on many fronts. The international manufacturers of construction machinery are responding to these challenges, and they will be presenting their innovations in all these fields at bauma 2013. bauma, the international trade fair for mining and construction machinery, mining vehicles, and building material machines, takes place from April 15 to 21, 2013 at the Messe Munchen exhibition center in Munich, Germany. The last bauma, in 2010, attracted 3,256 exhibitors from 53 countries, and more than 420,000 visitors from more than 200 nations. Covering a total of 555,000 m2 (5.97 million sq ft) of exhibition space, bauma is one of the largest trade shows in the world. This year?s fair is expected to surpass the 2010 figures. Here are some examples of the new developments visitors can expect to see at the show.

Jan 1, 2013

By John E. Auten

The computer program "Spike" is an "expert-rule based system" designed to fulfill two requirements. 1) acquaint unfamiliar personnel with big hole drilling techniques and 2) to stimulate the "old hands" in solving drilling problems. "Spike" will not make anyone a drilling expert The best way to understand a computer program is by example, but first some explanation of an? "expert-rule based system", As the name implies, the program tries to imitate a human expert. Unlike a normal computer program which uses a list of instructions to follow in a set order, a rule-base program has a collection of independent rules without an explicit list of things to do, one after the other. This type of program was mandatory to accommodate the ever changing situations and corrective actions that more accurately duplicate actual drilling operations and the advise of an experienced drilling expert.

Jan 1, 1988

By J. F. Hazzard, R. J. Chalaturnyk, C. Khazaei

"Various measurement, monitoring and verification (MMV) techniques are used to ensure the safety of operations in petroleum projects. One of these techniques is microseismic (MS) monitoring of the caprock. A challenge in MS monitoring of caprock is that unlike projects such as tunnelling in hard rocks, the number of recorded events are much smaller and also wide spread through the rock mass which makes interpretation of the physical meaning behind the events very difficult.In this paper, a numerical technique is introduced and validated for coupling a finite difference code, FLAC3D (Fast Lagrangian Analysis of Continua), with a discrete element code, PFC3D (Particle Flow Code). Then, a similar approach is adopted for modeling microseismic events in the caprock. For this purpose, the underburden, reservoir and overburden layers are modeled by FLAC3D. A rather small section of caprock above the reservoir is nulled in the FLAC3D model and substituted by a PFC3D inclusion. The reservoir pressures are then applied to the reservoir section of FLAC3D. Once the FLAC3D cycles, there will be induced pressures in the caprock due to its undrained response to the change in pressures within the reservoir. These induced pressures are transferred to PFC3D as boundary velocities. Then the PFC3D cycles and sends back the boundary forces to FLAC3D. This back and forth process continues until equilibrium is reached. During this process, there will be bond breakages within the PFC3D inclusion that can be regarded as MS events. An algorithm for recording the microseismic energies based on a change in local strains is also implemented. The results are then compared with real data. The effect of injection fluid temperature is also studied. No significant appearance of microseismic events was observed due to injection of CO2 in both cases (i.e. with or without considering temperature). However, this continuum/discontinuum approach seems promising for understanding the microseismic response of caprock in more detail and also using the MS data for validation of geomechanical models."

Jan 1, 2015

The Pike River Coalfield is located on the West Coast of the South Island of New Zealand near the top of the Paparoa Range approximately 45 km north east of the coastal town of Greymouth. The Pike River Coal Company Ltd (PRCC) is owned jointly by New Zealand Oil & Gas Limited (NZOG) and private investors.   The project has followed a long and difficult road towards project implementation which is now a reality with mine construction scheduled to commence in November 2005. The first stage of the project, the consent and land access process has been the most difficult. Resource consents for the project were first applied for, granted and appealed in 1998. A second round of modified resources consents followed a similar path in 2001. Following consultation with a range of parties including local and central government as well as the Department of Conservation, in March 2004 the Minister of Conservation granted PRCC access through Department of Conservation land. In October 2004, the Environment Court issued a Memorandum of Consent granting resource consents with conditions that had been negotiated between PRCC and the appellants.   The second stage of the project, investigation and design, commenced in December 2004 and was completed in June 2005 with the PRCC board ratifying the project capital and operating budgets. While the investigation and design work was proceeding, PRCC were also engaged in securing private equity and bank debt funding for the project. Formal Board approval to proceed with project construction was granted in September 2005 following completion of fund raising.   The project consists of a number of significant construction activities, namely: +        10.5 km of mine access road through previous logged and virgin rain forest; +        2.2 km of tunnel through metamorphic rock; +        a coal preparation plant about 8.3 km from the mine portal; +        an underground coal preparation plant; and +        a major slurry pump station, high and low pressure underground water supply systems to support hydraulic mining units and underground coal transport respectively.   Project construction is expected to be completed in April 2007 with completion of the ventilation shaft. Mine production will commence in December 2006 with coal from pit bottom development which will be carried out by the mine owner.

Jan 1, 2005

By AIME AIME

CHARLES A. CHASE, manager of the Shenandoah-Dives Syndicate, operating the Shenandoah Mines in southwestern Colorado, reviewed the current work at that property at a recent meeting of the Colorado Section, in Denver. This operation is a new one, having been in effect only two years, but the properties that were consolidated by Mr. Chase had had certain production in earlier years, since when they had: been inactive. These properties are located. in the high mountains of the San Juan section in Colorado, and although they had had a substantial production of ores of good grade in the upper levels, the project as a whole involved the bold undertaking of driving through the mountain, starting on the Mayflower claim on the Arrastre Gulch side and passing through the Mayflower, Slide, Terrible, North Star, Dives and Shenandoah No. 3 claims. This is a distance of over 8,000 ft. which is known to be mineralized, and the main operating tunnel was started at an elevation of approximately 11,000 ft. The main exploration work has already extended in nearly 4,200 ft. and the present mining production of nearly 550 tons daily is being secured while the main exploration work is being continued. The main heading is now within the North Star claim and the plan of operations involves continuing ahead for a few hundred feet and then starting an upraise which will he 2,200 ft. high and which will connect with the old workings of the North Star mine. The plan of ruining then will he to drive levels along the vein in both, directions to daylight and retreat downward, taking out all the ore and abandoning, the workings after the ore is removed. In this way one level after another progressing downward will be abandoned. The raise will be of two compartments, one of which will carry the 8-ton skip and the other the counterweight and manway. So far the ore has been broken over an average width of more than 11 ft., with an output of 39 tons per machine shift.

Jan 1, 1932

By Herbert Clark Hoover, Lou Henry Hoover

DIGGING of veins I have written of, and the timbering of shafts, tunnels, drifts, and other excavations, and the art of surveying. I will now speak first of all, of the iron tools with which veins and rocks are broken, then of the buckets into which the lumps of earth, rock, metal, and other excavated materials are thrown, in order that they may be drawn, conveyed, or carried out. Also, I will speak of the water vessels and drains, then of the machines of different kinds,1 and lastly of the maladies of miners. And while all these matters are being described accurately, many methods of work will be explained. There are certain iron tools which the miners designate by names of their own, and besides these, there are wedges, iron blocks, iron plates, hammers, crowbars, pikes, picks, hoes, and shovels. Of those which are especially referred to as " iron tools " there are four varieties, which are different from one another in length or thickness, but not in shape, for the upper end of all of them is broad and square, so that it can be struck by the

Jan 1, 1950

By Joseph S. Malesky

The discovery and development of explosives mark one of the most important findings in the history of civilization. Without explosives our vast economic enterprise concerning the mining of coal, copper, iron, aluminum, and various other ores could not function. Explosives are used in all kinds of construction works, in the razing of old buildings, blasting of trenches, driving of tunnels, and in many other instances common to our everyday life. For these reasons explosives are rightfully referred to as the forerunner of progress. An explosive may be defined as a substance (usually a mixture) that, when flame or shock is applied, yields readily to rapid combustion or oxidation, accompanied by the formation of such relatively large quantities of gases as to produce much violence and pressure which, in turn, react against their surroundings and cause what is termed an explosion. By their very nature the action of explosives is destructive and disruptive of adjacent materials. Decomposition of explosives can occur through deflagration, which is rapid burning or combustion, as in the case of black powder, or through detonation, which is an almost instantaneous disruption, as in the case of dynamites and blasting agents.

Jan 1, 1981

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

The integrated grouting method, as developed in the USSR, can be applied to multi-purpose operations that in¬clude the construction of shafts, drifts, and tunnels. In the USSR it has been used extensively for primary and final grouting of underground excavations in mining engineering and civil engineering projects. It has been applied in projects where the primary objective was waste contain¬ment. It has also been used to control subsidence beneath buildings located over mined out openings and to eliminate the flow of ground water beneath and around dams. This chapter considers several industrial applications in the form of case histories. The word integrated means that grouting activities were integrated with other activities. 9.1 CASE HISTORIES OF GROUTING FRACTURED ROCK WHEN SINKING VERTICAL SHAFTS The integrated grouting method has been applied most widely in the sinking of vertical shafts in fractured, satu¬rated rock. As explained previously herein, grouting oper¬ations preferably are carried out from the ground surface where they are integrated into the schedule of the setting up of the shaft excavation and construction equipment. Such integration reduces significantly the length of time required for preparation of shaft construction and the time required to sink the shaft. The increased efficiency is achieved by the elimination of cementing operations from the working face upon the penetration of each aquifer. Concomitantly, labor and energy consumption during shaft sinking are minimized because the more complicated and labor-intensive work is carried out at the surface section. Table 9.1 presents the essential parameters for the pri¬mary grouting of saturated fractured rock carried out by the integrated method when sinking vertical shafts. 9.1.1 NAGOL'CHANSK MINE NO. 1-2, VENTILATION SHAFT NO. 1 The integrated method of grouting saturated fractured rock through holes drilled from the surface was employed for the first time during the sinking of vertical shaft No. 1 at the Nagol'chansk mine No. 1-2 in the Don Basin. The shaft had an inside diameter of 6 m. The grouting operations commenced by using cement grout with a density of 1.7 to 1.8 g/cm3. Clay-based grout with cement and other addi¬tives was used only for grouting the aquifers where cement grout proved to be ineffective. By the time of the grouting of ventilation shaft No. 1, hydrodynamic analytical methods were well in hand and a preliminary method for designing isolation curtains had been developed. A method for designing and drilling ori¬ented directional drillholes whose natural curvature re¬flected the characteristics of the fracturing had been devel¬oped. The majority of the necessary equipment had been developed. Clay-based grouts were considered ready for the first industrial application. The hydrogeologic environment at the Nagol'chansk mine No. 1-2 contained many prolific aquifers. According to monitored drillhole data, the total expected inflow into ventilation shaft No. 1 without grouting was 425 m3/hr. This large ground water inflow rate was supported by in¬formation obtained during the sinking of the main shaft and from the auxiliary shaft of the "Nagol'chansk" mine No. 1-2. The construction contractor at the site began sinking ventilation shaft No. 1 to a depth of 217 m by cement grouting from the working face of the shaft. As a result of the cement grouting operations, the inflow of water was reduced from 105.4 to 40 m3/hr over this 217 m interval. However the residual inflow rate caused the shaft to become functional at the depth of 217 m. At this point all grouting operations were transferred to SPETSTAMPONAZHGEOLOGIA (STG) in an effort to improve the operation. In order to carry out preliminary grouting from the sur¬face, seven drillholes were installed and designed for re¬ceiving grout. Drilling was implemented by the ZIF-1200A rig to the design depth of the shaft. The geometry of the principal fracture system that would be encountered by the shaft was estimated while breaking ground using informa¬tion from other shafts. The drillholes were arranged so that they cut through the aquifers uniformly around the shaft. Some preference was given to intercepting the aquifers that were located along the rock dip above the shaft bottom because it was established that the ground water in the aqui¬fers flowed downward along the dip into the cone of de¬pression of the shaft. Investigations in the drillholes using the DAU-3M flowmeter described in Chapter 3 showed that all of the nine aquifers were penetrated by the time a depth of 690 m had been reached. The flowmetric information showed that aquifer numbers [1 and 2], [4, 5, and 6], and [7 and 8] had practically identical hydrogeologic characteristics; so it was decided to connect them into combined stopes and inject the grout using five stopes instead of nine. Table 9.2 shows the results of the grout injection into the aquifers. The volume of grout pumped into the separate intervals through single holes varied from 204 to 15 m3. The trend of grout consumption showed a reduction of grout consump

Jan 1, 1993