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By D. S. Cavers

INTRODUCTION This paper discusses aspects of rock engineering applicable to surface gold mines. Discussion particularly centers on areas in which rock mechanics can enhance the mine productivity or profit- ability. Methods of analysis are discussed briefly but for more detai1s the reader is referred to the appropriate references. A team approach to the overall design of the pit generally pro- duces a better result at an overall lower cost. In particular, money is wasted by asking for a "report on slopes" and then input- ting this slope angle into a pit design and economic analysis study without further reference to the geotechnical engineer. Pit slope angle depends on slope orientation, location, pit shape, pre-existing discontinuities and mining methods. Preliminary slope angles can be produced but optimization depends on ongoing inter- action between the mining engineers, geologists and geotechnical engineers. This interaction can a1 so substantially reduce overall design time which may be critical in bringing remote plays into production. Most surface gold mines in North America are in igneous or metamorphic environments. Rock mechanics aspects discussed in this paper will be discussed with reference to these environments al- though most of the principles and comments are applicable to sedimentary environments as we1 1.

Jan 1, 1987

By Lynn I. Kelley

Industrial-mineral production contributes on the order of CDN$1.5 billion per year to the economy of Saskatchewan, Canada. Potash is the primary industrial mineral produced in the province, followed by aggregate, sodium sulphate, salt, potassium sulphate, peat, clays, silica sand, calcium-chloride brine, and "clinker." Potentially economic deposits of kaolin, diamond, and building stone are under evaluation. All of the major deposits of commodities discussed here are located south of lat 56°30'N. Most of the locations are presented on Figure 1 POTASH Saskatchewan supplies approximately 30% of the world's demand for potash. Eight conventional underground mines and two solution mines produced an all-time high of more than 13 million metric tonnes of potash (KC1) (8 million tonnes K2O equivalent) in 1997, valued at over CDN$1.4 billion. All of the mines exploit the middle Devonian Prairie Evaporite. The four potash-bearing members of the Prairie Evaporite underlie much of southern Saskatchewan and extend southward into North Dakota and Montana and east- ward into Manitoba (Fig. 2). Halite and anhydrite separate the potassic members. The evaporite beds locally are interrupted by solution structures (anomalous ground). Some of the structures may be related to deep-seated tectonic features.

Jan 1, 1999

By John A. Dearinger, Fred D. Wright, Robert C. Howell

INTRODUCTION The rock mechanics study of the shortwall mining system at Beth-Elkhorn corporation's Hendrix No. 22 Mine is being conducted under contract with the U.S. Bureau of Mines. The objective of the study is to obtain data on surface subsidence over a portion of the mined area, strata subsidence in the mined area, and underground stress changes and displacements in and near natural support elements of the mining system. This study is currently in progress, and only a portion of the results are being presented at this time. The Hendrix No. 22 Mine is located in the moderately rough mountainous terrain of Letcher County in eastern Kentucky. The nearest mine workings to the shortwall area are approximately 1,900 feet distant. The depth of cover over the mined shortwall area varies from 770 feet at Panel 1 to 170 feet at Panel 5 (Figure 1). The strata in the shortwall area is summarized in the three core logs shown in Figure 2. Hole no. 2 is located in approximately the center of the mined shortwall area (Panel 3, Line C, Figure 1). The coal seam in the shortwall area is approximately 4 feet thick and horizontal, although one major ‘roll’ intersects all panels.

Jan 1, 1977

By Joel K. Waterland

The Homes take Mining Company at Lead, South Dakota has a yearly production of 1,868,741 tons of ore produced from horizontal cut-and-fill stopes and timber cut-and-fill stopes. This production, in addition to the necessary development (drifts and raises), requires 5,168,271 feet of drilling annually. The rock for the most part consists of quartz; mica schist; and cummingtonite schist; and must be considered as relatively hard. Penetration rates average about 15 inches per minute using an operating air pressure of 100 psi. The rock is tough, rather than brittle, and dulling occurs quite rapidly. The average regrind period is approximately 19 feet per regrind using a 1-7/16 inch carbide bit. At the present time, all drilling is accomplished with 2-5/8 inch bore airleg mounted drills, and 2-3/4 inch and 3-inch diameter bore -stopers (raising only). These machines are all direct striking, with 4-1/4 inch chucks. 7/8-inch hexagonal carbon drill steel with forged collars are used in conjunction with 1-7/16 inch multi-use, threaded, carbide bits. 62,793 rods, and 16,234 bits are used to maintain the yearly production. The number of rods noted are the rods that were sent: underground, both new and reconditioned.

Jan 1, 1963

By Y. C. Kim, M Nataraja, J. T. Buckley, D. H. Tiktinsk

The Nuclear Waste Policy Act of 1982 (NWPA) assigns the U.S. Department of Energy (DOE) the responsibi1it.y of locating, constructing. operating, closing and decommissioning a nuclear waste repository. Prior to submittal of a licensee application to the U. S. Nuclear Regulatory Commission, the DOE is required by 10 CFR part 60 to conduct a program of site characterization for the site to be described in such application. Site characterization includes an exploration and research program in the field and laboratory to determine the geologic conditions and ranges of those parameters which are necessary for determining the suitabi1it.y of the site as a geologic repository. The information thus collected should be able to establish with reasonable assurance that the public and environment will be adequately protected from the hazards associated with a repository at the site. This paper demonstrates by example a geostatistical approach which, on the basis of the existing limited data base, can be used to evaluate (1) whether existing information on a certain parameter is adequate to be considered representative. (2) if additional information is required, where should the additional sampling or testing be performed, and (3) whether the new1.y acquired data along with the existing information constitute representative data.

Jan 1, 1986

By Charles C. Y. Chen, Duane R. Skidmore

Batch leaching of Ohio #1, Ohio #4A and Ohio #5 coals was conducted with the thermophilic. sulfur, oxidizing microorganism: Sulfolobus acidocaldarius. Four different size fractions of each coal, with median sizes of 508, 200, 90, and 49 microns, were subjected for 10 days to leaching in a 5% coal-water slurry at 72°C and a pH of 2. Release rates and extent of sulfate and iron I1 and I11 ions from the coals during the leaching process and distribution of sulfur forms (sulfate, pyritic and organic) in the coal samples before and after treatment were determined. Data indicated that Ohio #1 coal, with an initial sulfur content of 0.6%, showed no significant changes with leaching in total sulfur content or forms of sulfur. The Ohio #4A coal, with initial total sulfur content near 5%, yielded product, for sizes below 200 microns with 90% reduction of pyritic sulfur while the extent of organic sulfur removal varied with coal particle size. Results similar to those for Ohio #4A were observed on the Ohio #5 coal, which had an initial total sulfur content of 3% and a final sulfur content near 1%. The findings were interpreted in terms of basic biological, chemical and physical mechanisms.

Jan 1, 1989

By J-Y Chen

China was once deficient in nickel and almost totally reliant on imports for its domestic needs, supplies mainly coming from Canada since the ending of imports from the U.S.S.R. in early 1960's. However, the discovery of the huge nickel/copper sulphide deposit in Jin Chuan, Gansu Province of northwestern China in late 1950's and its development and construction through the 1960's and 19701s, have much improved China's domestic nickel position. In 1985. China's nickel production has exceeded 27.000 tons, compared with only 7,000 tons in 1977. Today, Jin Chuan Nonferrous Metals Company operates two underground mines and a small open- pit; owns China's largest reserves of nickel and platinum, and the second largest reserves of copper and cobalt; and has become the most important subsidiary of China National Nonferrous Metals Industry Corporation (CNNC) which was established in April 1983. The Jin Chuan nickel-copper mines are located in Jin Chang City, Gansu Province, 1,700 km west of Beijing in the People's Republic of China (Figure 1). It is believed that Jin Chuan mines have the second largest nickel reserves in the world, after Sudbury basin, Canada. The deposits are situated on the north slope of the Longshu Mountains, 50 km north of Jin Chang City, and border on the southmost edge of the Gobi Desert plateau. The deposits were discovered in 1958. The total strike length of mineralized zones was approximately 6.5 km long, which was divided into four mining areas by faults (Figure 2). The Longshu underground mine of No. 1 mining area began development in 1959, with full ore production capacity of 1,000 TPD. The open-pit, also in No. 1 mining area, which was brought into operation in 1966 with ore output of 5,500 TPD, was closed down in 1982. The East Extension of the open-pit was put into operation in 1983; present production capacity is about 2,200 TPD. Nickel production of Jin Chuan mines in recent years, and production capacity of each mining operation, are shown in Tables 1 and 2.

Jan 1, 1987

"Use of the Bond Work Index to size industrial grinding equipment was widely known. Bond also described its application for measuring grinding circuit efficiency. Bond Work Index efficiency has been widely used by industry, but until now without concise standard methods, or a formal guideline, for doing so. The Bond Efficiency Sub-Committee of the Industrial Comminution Efficiency work group of GMSG has produced such a guideline, which follows. It covers the methods and provides examples of how to calculate the Bond Work Index Efficiency of most industrial grinding circuits. Bond Work Index testing equipment and methods are also provided. The next steps are: (1) to establish standard reference testing materials and one or more laboratories to serve as bases for reference testing; (2) to establish the accuracy of relative circuit Work Index Efficiency determinations; and (3), to develop and provide a public data base of industrial comminution circuit Bond Work Index Efficiencies. OUTLINE Introduction and Background Part 1. Determining the Bond Efficiency of a Grinding Circuit Method Demonstration/Example Calculations Bond Work Index Testing Equipment and Procedures Accuracy of Comparative Circuit WI Efficiency Determinations (under development) List of Symbols Part 2. Data Base of Bond Grinding Circuit Efficiencies (under development) References Annexes: A1. The Bond Crushing WI Test Equipment and Procedure for Bond Efficiency Determinations (Revised March 26, 2015) B1. The Bond Rod Mill WI Test Equipment and Procedure (Revised March 26, 2015) C1. The Bond Ball Mill WI Test Equipment and Procedure (Revised March 26, 2015) Introduction and Background Please see the references and slide presentation “Bond Efficiency SLC (rev3)-print.pdf.” This efficiency determination applies to most brittle materials in their naturally occurring (un-scalped) size distributions being treated in size reduction circuits down to a product sizing of approximately eighty percent passing 70 µm (see below). Unusually shaped materials (e.g. mica) should be regarded with caution. Part 1. Determining the Bond Efficiency of a Grinding Circuit Method In the plant: • Define the circuit for which the WI efficiency is to be determined • Procure samples of the circuit feed and product • Obtain the power draw of the size reduction equipment at the drive pinion(s) • Obtain the circuit throughput rate (dry tonnage) In the laboratory:"

Jan 1, 2016

By Robert H. Merrill

INTRODUCTION In most cases, the design of an underground mine is based upon the premise that the ground either will cave or will be stable. This chapter concerns the design of a mine in ground that will cave readily or with some as¬sistance, such as by long-hole drilling and blasting. Some of the more widely used caving systems of mining are panel caving, block caving, sublevel caving, and large pillar recovery. Some of the less widely used systems are glory-hole, top slicing, and induction caving. Al¬though the common practice of pillar robbing is not usually considered to be a caving system, this subject will be treated as a part of this chapter. BASICS OF CAVING Caving systems are most successful in ground that will cave in sizes that will flow through openings and grizzlies, and will easily load in cars or on belts for haul¬age. The ground most likely to cave well is highly frac¬tured and contains breaks, flaws, or other discontinui¬ties that form planes of weakness. Also, caving action can be greatly enhanced if the host rock itself is low in compressive, shear, and tensile strength. Ideally, a cav¬ing system of mining is best employed when the criteria for caving is a feature of the ore body and the develop¬ment drifts, haulageways, and drawpoints can be mined in a highly competent rock beneath the mineralized zone. However, the development is often in the same, or similar, fractured rock and the openings require sub¬stantial artificial support to assure stability. Several clues can be assembled to identify potential caving ground; however, for borderline cases, no sure method has been devised to date. The diamond-drill cores taken for exploration can provide an excellent clue provided drilling is performed carefully by experienced drillers. For example, if the ground is cored in such a manner that the breaks in the core are caused more by failure of the rock than by whipping core barrels, plugged drill bits, or other drilling causes, and the intact core lengths are consistently long [say, 0.6 to 3 m (2 to 10 ft) of unbroken core], there is little reason to believe the ground will cave without considerable as¬sistance. This is especially true for rocks with compres¬sive strengths above 34.5 MPa (5000 psi) and tensile strengths above 2.1 MPa (300 psi). On the other hand, if core recovery is low (below 80%) and the recovered ore is broken in small pieces and the breaks are along obvious weaknesses in the rock, the chances are excel¬lent that the ground will cave. This is true even when the rock between the defects has high compressive and tensile strength. Another clue has already been mentioned, that is, the measurement of the physical properties of the rock and the natural planes of weakness or defects in the rock. The planes of weakness in the rock can often be detected from outcrops, cores, or other exposures of the rock under consideration. Some rock types are known to be strong and will sustain large, unsupported open¬ings and would be difficult to cave intentionally. Yet the same rock type can also contain unbonded or weak planes of weakness or fractures, and in these locations the rock would undoubtedly cave with little assistance. Therefore, although the inherent strength of the rock is a factor in caving, the natural defects in the rock are more often the deciding factor. DESIGN CONCEPTS For the most part, the design of openings for caving ground is a problem of the interaction of openings over a relatively large area of the mine. To illustrate, Fig. 1 is a simplified section of a series of openings along the grizzly level or draw level of a block caving or panel caving development, and above this opening is a simpli¬fied section of a room-and-pillar arrangement on the undercut level. At this stage of the development, the stresses around the openings on the grizzly level are only moderately influenced by the openings on the undercut level and vice versa. Therefore, the stresses around the openings are approximated by the stresses around single or multiple openings in rock, the values of which are de¬scribed in the literature (Obert, Duvall, and Merrill, 1960; Obert and Duvall, 1967). Once the pillars on the undercut level are blasted (Fig. 2), the situation changes abruptly. The undercut opening (prior to caving) now can be approximated as an ovaloidal opening above the grizzly drifts and this opening tends to shield the vertical stress field. As the caved stage is drawn the stope approximates a much larger rectangular or square opening filled with rock, and if the rock is not sustaining a major portion of the stress field, this opening can be considered (for en¬gineering purposes) to be empty and the stresses that interact between the larger and the smaller openings take on a totally new perspective (see Fig. 3). Next, let the material cave to the surface, and let the caving ma¬terial sustain some stress, but much less than if the ma¬terial were intact. This condition is similar to a soft inclusion in a rigid body and has been treated in the literature (for example, Donnell, 1941). At this point in time, the grizzly drifts are subjected to the stress con-

Jan 1, 1982

By Timothy Batchler, Khaled Mohamed, Morgan Sears

The National Institute for Occupational Safety and Health (NIOSH) initiated a testing program to assess the effects of lateral loading on various standing supports used in underground coal mines. This study is not intended to endorse a particular type of standing supports nor to provide a comparison between different types of standing supports. This paper aims to measure the lateral loading capacity of two types of standing supports and evaluate how it affects their vertical loading capacity and yielding mechanisms. The study utilizes the Mine Roof Simulator (MRS) at NIOSH Pittsburgh. Type-1 support is characterized by telescopic tubes made from ASTM A-513 steel with a specified locking mechanism, and Type-2 support comprises a single component constructed from ASTM A500 Welded Grade C pipe. Vertical load tests were carried out to assess the designed performance of Type-1 and Type-2 props. Type-1 props demonstrated a "constant yielding" behavior, achieving an ultimate capacity of up to 140 kips. Type-2 props showed a "load shedding" behavior, reaching peak verti- cal loads of up to 191 kips. Buckling analysis revealed that Type-1 props were not designed to buckle, and Type-2 props failed due to buck- ling rather than material yielding. Under vertical pre-loading, lateral loading tests showed that Type-1 props can withstand about 9 kips, and Type-2 props can withstand about 10.79 kips of lateral load. This represents roughly 6% of their vertical loading capacities. Testing props under lateral loads in addition to their intended vertical loads showed that the lateral loads could influence their loading mechanisms. Type-1 props did not exhibit the expected "constant yielding" behavior under lateral load but instead showed "load shedding" decreasing their rated vertical capacity by 70%. At high vertical pre-loading, Type-2 props demonstrated "constant yielding" behavior instead of the anticipated "load shedding," leading to a 60% reduction in their rated vertical capacity. These findings provide valuable insights into the behavior of these props under different loading conditions, aiding in their efficient use and design in underground mining operations.

Jun 25, 2024

By Robert G. Lundquist

REAM (for Rapid Excavation And Mining) is a method of rock fragmentation by high-velocity projectile impact developed by Physics International Company. This paper reports the results of field testing over the last two years. The success of the method is based on the fact that the amount of rock broken per shot scales nonlinearly with projectile impact energy. Thus, high energy impact yields spectacularly low specific energies for rock breakage. The field work reported used 90- and 105-mm military guns to accelerate 8-1/2 and 10-pound concrete and steel projectiles to velocities of 5000 to 5500 fps. All tests were performed in a competent 25,000 psi granodiorite at Hope Valley, California. During the first year of field testing 26 feet of 13-foot-diameter tunnel were produced with the 105-mm gun. The first 20 feet of this tunnel were line-drilled to aid in producing a stable portal. During the second year the tunnel was extended an additional 29 feet using the 90-mm gun. The 55-foot tunnel was produced by 590 shots. Between 1-1/4 and 1-1/2 tons of granite were broken by each shot. A silenced 90-mm gun was placed underground and the tunnel portal closed with a steel barrier. Overpressures of less than 1 psi were measured on the inside of the barrier, corresponding to an impulse noise of about 170 db. The muck produced was similar to drill and blast muck. Overbreak and ground support problems appeared to be less than with drill and blast. The contour of the tunnel was easily controlled using the 90-mm gun; the tests were also conducted using a 57-mm gun for trimming. Projections indicate that advance rates with REAM will be three to seven times faster than drill and blast, at less than two-thirds the cost. This is based on the development of a fully integrated continuous system and an inexpensive propellant giving a total expendables cost of less than $3 per shot. Other applications of the REAM technology are found in secondary breakage for open pit mining, in drilling large diameter holes, and in underground mining for longwall mining, overhand stoping, and raising.

Jan 1, 1974

By J. Herbst, J. A. Kitchner, J. P. Anfruns, G. Poling

Measurements have been made of the rate of capture of single particles of strongly hydrophobic, surface-methylated, quartz, of sizes ranging from 12 to 40 pm diam., by single rising bubbles of diam. 0.5 to 1.1 mm. The results agree closely with the theoretically calculated hydrodynamic collision rates, indicating that, in this system, every collision results in capture and retention of a particle.

Jan 1, 1976

By Steve Chorley

Open-type, or main beam, tunnel boring machines (TBMs) can achieve impressive rates of production. This type of TBM typically outperforms other types of TBMs currently in use around the world. In good ground conditions, main beam machines have achieved production rates of more than 1 km/month (0.6 miles/month). However, installing effective ground support in poorer rock conditions on main beam machines can be a major cause of concern. The design of main beam machines is such that installation of ground support equipment, particularly at the front of the machine in what is commonly known as the L1 support zone, has traditionally been limited to rock drills and basic working platforms. Personnel working in the L1 zone have, until recently, been protected by roof supports and extended fingers, often with restricted access to working areas.

Jan 1, 2011

By Kerry Gee, Gary Colgan, James Fricke

Introduction Elevated zinc concentrations are observed in groundwater within alluvium adjacent to Silver Creek east of Park City, Utah. The study area is located in a small valley one and one-half miles northeast of Park City, Utah in Summit County, Utah (Figure 1). Although surface water concentrations in Silver Creek are typically less than 1 mg/l, groundwater concentrations have been measured as high as 85 mg/l in a near neutral pH environment. The elevated zinc concentrations in groundwater result from several factors including: (1) dissolution of zinc from tailings exposed at the surface and within the alluvium, (2) oxidizing conditions created by surface/ground-water exchange, and (3) ideal geochemical conditions for the dissolution of zinc-containing minerals. These ideal conditions include readily soluble minerals, available anions for complexation, and lack of available surface adsorption sites.

Jan 1, 2005

By Michael Feroz

Severe drought in the Colorado River Basin over the past 10 years has caused water levels to drop in Lake Mead by more than 113 feet. The current lake level is at elevation 1,112 feet mean sea level (msl). Lake Mead is the primary source of water supply for greater Las Vegas. Water for Las Vegas from Lake Mead is accessed through two existing submerged intake shaft and tunnel systems. If the water levels continue to drop in the lake, the Intake No. 1 will become inoperable below water level elevation of 1,050 feet msl. To sustain the current levels of water delivery to all member agen-cies in the southern Nevada region, the Southern Nevada Water Authority (SNWA) has decided to construct a third intake at a deeper depth in the lake to ensure better quality water and enhanced reliability.

By M. D. Johnson

Mina Santa Rosa, in west central Panama, is the first large-scale gold mine in Panama and will soon be the largest single-mine gold producer in central America (Fig. 1). At full production, the mine will process 1.8 Mt/a (2 million stpy) of ore and produce more than 1.8 t/a (60,000 oz/year) of gold and 9.3 t/a(300,000 oz/year) of silver. Mina Santa Rosa is a subsidiary of Greenstone Resources of Toronto, Ontario. Freeport-McMoRan geologists first visited the area in 1985. Their rock chip sampling adjacent to old shafts and adits confirmed widespread gold and trace element geochemistry anomalies. Surface mineralization appeared to be localized in and adjacent to narrow veinlets. However, they recognized astructural setting similar to the Nevada Tertiary volcanic-hosted epithermal gold systems.

Jan 1, 1996

By George A. Desborough

Marine black shales of Paleozoic age in the Western United States that exceed 10 weight percent organic carbon contain as much as 5,000 ppm vanadium, 1,000 ppm molybdenum, 1,300 ppm nickel, 600 ppm cobalt, 16 ppm silver, 7,000 ppm chromium, 18,000 ppm zinc, 280 ppm cadmium, and 350 ppm selenium. Depositional settings include: (1) continental-rise and marginal-ocean basins, (2) continental-shelf and foreland basins, and (3) cratonic-platform embayments. Most of these metalliferous organic-rich shales have not been examined for resources because of rapid weathering and poor exposures which preclude easy sampling of unoxidized rock necessary to determine actual concentration and residence of indigenous metals. Generally, the marine black shales of Paleozoic age in the Western United States are richer in metals than marine Paleozoic black shales in the Eastern United States.

Jan 1, 1982

By W. E. Hale

The Institute of Applied Geology (I.A.G.), a joint project sponsored by the University of the Philippines, the Philippine Bureau of Mines and the United Nations Development Programme (Special Fund) undertook during 1965-66 a number of field projects in Northern Luzon Island (Fig. 1). These field operations were of a two-fold nature: during approximately six weeks an orientation survey, involving, geology, geophysical surveying, and geochemical prospecting, was conducted over a partly developed, disseminated copper deposit at Sto. Nino (Fig. 1). The Sto. Nino study demonstrated (Hale and Govett, 19677 Govett and Hale, 1967) the feasibility of geochemical prospecting relative to other exploration methods for the location and delineation of disseminated copper occurrences, the major source of metal in the expanding mineral industry of the Philippines. A vein occurrence at Suyoc (Fig. 1) was similarly subjected to

Jan 1, 1968

By W. S. Sanner

Experiments were conducted to determine the quantity and purity of ultraclean anthracite that could be prepared in the laboratory, using conventional separating techniques. A low-, a medium-, and a high-volatile anthracite were studied. Specific gravity separation followed by grinding, screening, and refloating at lighter specific gravities yielded small quantities of material purified to 1.0 percent ash, as follows: From low-volatile anthracite, 0.7 percent; from medium-volatile coal, 1.0 percent; and from high-volatile coal, 2.3 percent. The reduction in ash had no appreciable effect on the volatile matter or sulfur content of the coal. The limited quantities of material containing 1.0 percent or less ash obtainable indicate that if ultraclean coal is desired, a practical approach should include the processing units as an integral part of a conventional dense-media washing plant.

Jan 1, 1968

By R. L. Beatty

The Bond International Gold (BIG) Colosseum Gold Mine was discovered in 1865, but no recorded production occurred until the 1930's, with production approximately 19k9 (615 oz) of gold. The mine was closed in 1942 as a non-essential industry during World War II. The property was acquired by Dallhold Resources in mid 1986. Construction of a 1.1 million Mt carbon-in-pulp, cyanide mill started in November, 1986, and was completed in September, 1987. Minable reserves are estimated to be 9,560,831Mt (10,539,000St), with an average grade of 2.1 g/Mt (0.062 oz/st). This paper describes a general view of process flow and plant equipment, as well as a look at the maintenance and operational problems encountered in the fist six months of operation.

Jan 1, 1989