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By Steve Dillingham

Today, a typical blaster is likely to wear many hats, including, but not limited to, supervisor, risk manager, safety coordinator, explosives engineer, communicator, as well as professional decision maker. Simply put, a blaster in the new millennium is an all encompassing blast site manager. As you know, accidents on the blast site can range from blast crew injuries to blasting incidents resulting from vibration, flyrock, misfires, etc. Not surprising, the incidents can quite often be traced back to direct preventable human action on the blast site. They can, however, frequently be avoided, if responsibly identified early and managed properly by an efficient blast site manager. Your company cannot afford to simply employ blasters; good blast site managers are necessary to produce the safest results possible.

Jan 1, 2001

By Benjamin C. Pollard

The location and production potential of a large block of strippable reserves in North Dakota were determined by using published data as a base and adding new drill hole data or other data contributed by companies that presently own or lease coal lands. Only beds exceeding 5 feet in thickness, under less than 120 feet of overburden, and in large blocks of 5 million tons or more were included in the estimates. All such reserves are in the Fort Union Formation of western North Dakota and are lignite in rank. Sixteen large blocks of reserves evaluated in this study are estimated to contain a total of 4.1 billion tons of strippable lignite.

Jan 1, 1972

By A. C. Johnson

The T. L. shaft of the Eureka Corp., Ltd., near Adams Hill, Eureka, Nev., was sunk in 1954 and 1955 to permit exploration and development of the lead-zinc-silver ore that had been indicated by drilling. This paper describes the sinking of the 4-compartment, vertical shaft to a depth of 1,127 feet and includes pertinent details of equipment, methods, and costs. A shaft crew of 10 men, supplemented by 13 hoistmen, topmen, carpenter, and shop crewmen, sank and timbered the 1,127 feet of shaft in 630 workdays. Total labor per foot of shaft was 107.95 man-hours. Supplies per foot of shaft included 20 pounds of explosives, 1.61 pounds of drill steel, and 200 board-feet of timber. Power consumption, including that required for pumping, averaged 3,726 kw.-hr. per foot.

Jan 1, 1958

By C. H. Johnson

The Big Jim mine of the United Eastern Mining Co. is at Oatman, Mohave County, Ariz. Oatman is in the foothills on the western side of the Black Mountains, at an elevation of about 3,000 feet. It is on a paved highway about 30 miles west of Kingman, the principal distributing; center for mining supplies in the district, and 25 miles from Topock, en the Colorado River. Both latter towns are stations on the Santa Fe Railroad; Oatman has no roil connection. Mining conditions are typical of western Arizona; water is scarce, no mine timber is available locally, and the winters arc mild and the summers hot. The supply of a killed labor is usually adequate, but in 1933-34 it w not more than sufficient for the restricted operations. Wages have been $4.00 to $4.50 for miners. Electric power was supplied from Kingman and cost about 2.6 cents per kilowatt-hour.

Jan 1, 1935

By W. G. Pariseau

Conventional barrier pillars in underground coal mines are intended to protect mains and sub-main entries for relatively long periods of time. Inter-panel barrier pillars that may be used in especially deep coal mines, say, below 800 m (2,400 ft), are intended to protect longwall panel faces and adjacent panel entries for the life of nearby panels. Barrier pillar failures have serious consequences, especially when failure is sudden and catastrophic. Design of barrier pillars is thus of considerable importance to safety and efficiency of operations. Locally, failure occurs at the limit to elasticity when stress tends to exceed strength. Strength of rock is stress dependent, so an analysis of stress is central to rational design of barrier pillars. There are several approaches to stress analysis for coal mine pillar design that range from the simplified but popular tributary area method including variations that use empirical definitions of the tributary area (volume) to more sophisticated numerical methods such as the boundary element method and the finite element method. Despite ease of use, empirical methods are limited and always suspect when applied to conditions beyond their scope of definition, while boundary element methods do not generally include pillars in the solution domain. The finite element method is not so restricted and thus is a natural choice for conventional and inter-panel barrier pillar safety analyses. Caving of roof strata certainly affects the distribution of stress about longwall faces and nearby barrier pillars. Conventional wisdom that considers roof strata to respond as cantilever beams indicates the longer the span the greater are loads on longwall faces and face supports. Failure of roof beams, caving, shortens spans and relieves face and support loads. The same view is applicable to barrier pillars in that long roof spans outside of barrier pillars leads to higher pillar loads than short spans. Hence, caving should be beneficial to barrier pillar safety. However, the benefit may be small depending on the extent of caving and only in the form of a slight reduction in overburden weight transmitted to the barrier pillar, especially in case of inter-panel barrier pillars. FINITE ELEMENT MODEL Finite element models for stress analysis in rock mechanics are formulated in consideration of physical laws, kinematics, and material laws. These models meet the essential requirements for equilibrium of stress, satisfy strain displacements relations, and in a general sense, relate stress to strain through material laws, often referred to as constitutive relations. The last may be complex and replete with nonlinearities. A de facto engineering material law is the elastic model in the form of Hooke?s law for linear, homogeneous materials that may be anisotropic. A limit to elasticity is imposed by strength in the form of a failure criterion or yield function that accommodates triaxial stress states. Examples used in rock mechanics are the well-known Mohr-Coulomb (MC)criterion, the popular Hoek-Brown (HB) criterion, and the Drucker-Prager (DP) criterion. The first is linear in stress, while the second is nonlinear. Both do not include any effect of the intermediate principal stress. The last is also linear in stress but includes the intermediate principal stress. All show an increase in the major principal stress at failure with confining pressure. In the UT2/3 finite element programs, the failure criterion Y is a nonlinear, anisotropic form and is also the plastic potential [1]: [YJI=+ 11NYJI=+- (1) ] where N is an exponent, usually 2, and J2 and I1 are nonlinear anisotropic forms of the second invariant of deviatoric stress and the first invariant of principal stress. The form (1) reduces to DP when N=1and isotropy prevails. When Y = 1, the limit to elasticity is reached and plastic deformation is possible. The material constants that enter (1) are all computed from unconfined compressive, tensile, and, in case of anisotropy, shear strengths. These latter strengths may be estimated from conventional laboratory test data. Of course, joints in field-scale rock masses induce different values. How different is often a matter of conjecture. For comparative studies, laboratory rock properties seem adequate. Inelastic, that is, plastic behavior occurs with strain beyond the elastic limit. Strain hardening, ideal plasticity, and strain softening are possibilities. Ideally brittle behavior with complete loss of strength is an extreme form of strain softening [2]. Caving may also be considered as strain softening with a less drastic reduction of material properties. As intact roof strata fail, fall, and form a rubble pile below, a new material is formed with greatly reduced elastic moduli and strengths, say, 1/100th of the original intact moduli and strengths. From another viewpoint, caving is simply a natural cut and fill process. Input for stress analysis by the finite element method includes: 1. Preexcavation stress state 2. Geology (stratigraphic column) 3. Structure (major faults, joint sets) 4. Elastic moduli 5. Strengths, and specific gravities of each rock type 6. Excavation sequence Output from an analysis includes distributions of: 1. Stress 2. Strain 3. Displacement 4. Local factor of safety (element fs) The last is defined as the ratio of strength to stress and in view of the failure criterion (1), is the ratio [fsJstrengthJstress=] where the numerator is the maximum possible value and the denominator is the actual value of the square root of the second invariant of deviatoric stress. This invariant is a measure of shear stress intensity but also reduces to uniaxial tensile and compressive strengths as the case may be. An fs greater than 1 indicates an elastic state of element stress; a value of 1 indicates the limit to

Jan 1, 2009

By Herbert L. Lines

This paper is one of a series on mining and milling methods and costs published by the Bureau of Mines. The Ash Peak mine of the Veta Mines, Inc., Duncan, Ariz., is of particular interest in that the only metal of commercial value in the ore is silver. The silver is concentrated by flotation; the concentrate is shipped to a smelter for treatment.

Jan 1, 1940

By J. J. Taber

This is the final report on Grant No. G0144067. The report presents the results of a number of laboratory studies carried cut in several different ways to determine the effectiveness of foam for blocking the flow of gas in coal. The effectiveness; (of foam for retarding or blocking the flow gas in various porous media has been observed by many investigators1-4 but we are not aware of similar detailed studies in coal. Because of the diversity of our experiments, the key findings and general conclusions are presented very briefly in Section II. As a convenience to the reader a complete table of cements is included to enable the reader to find those experiments and studies of greatest interest.

Jan 1, 1976

By Robert A. MacGregor

"Following several months of test work, water leaching in mined-out areas at Stanrock commenced on a production basis in February, 1963. A combination of fast high-pres-sure washing and rewashing of stopes on a predetermined cycle, and the intermittent sprinkling of low-grade heaps, using both fresh water and recirculated acid water, has proved most successful. There was also a correlation of daily sampling data points to a relationship between the leach water grade and the grade of the material being washed, the leaching time and the location along the dip. An increasing tolerance for uranium by the acid-producing bacteria is also indicated. A steady production rate of 10,000 to 12,000 lbs. of u.08 per month is being maintained, with an over-all mine water grade of 0.25 lb. per ton and a pH of 2.3."

Jan 1, 1966

By B Beamish

The concept of a ‘critical velocity zone’ in the longwall goaf environment for the development of a spontaneous combustion event has been supported by numerical modelling. However, there is no experimental data available for Australian coals to show ventilation effects on the self-heating incubation behaviour of broken coal in the longwall goaf. A preliminary study using incubation testing has been completed on three high volatile bituminous coals (A, B and C) with R70 self-heating rate values of 3.86°C/h, 3.15°C/h and 2.09°C/h and moisture contents of 11.3 per cent, 3.9 per cent and 4.7 per cent, respectively. Mass to flow rates used in the adiabatic oven reaction vessel have been selected to be indicative of a sluggish, natural air leakage and medium ventilation environment. Under a sluggish ventilation flow rate all three coals incubate to thermal runaway in minimum time frames of 2.8–3.1 years, 34–42 days and 79–95 days, respectively. When the ventilation flow rate is increased to natural air leakage, coal A with the highest R70 value and highest moisture content selfheats to 28°C above mine ambient temperature before heat loss takes over due to moisture evaporation. However, both Coals B and C at this flow rate incubate to thermal runaway in a shorter time frame than under sluggish flow. At the medium ventilation flow rate, only Coal B incubates to thermal runaway, albeit after a significant extended period of time. These results show how important it is to determine the heat balance between the coal intrinsic reactivity (as indicated by the R70 value) and moisture content. The results also have implications for the possible location of the ‘critical velocity zone’ and subsequent hotspot development and migration.

Oct 12, 2022

By Lloyd A. Morley

Two methods are presented to determine the mechanical shaft torque as a function of time for an induction motor. The methods are based on the motor equivalent circuit and use as input variables of the motor terminal voltage, line current, input power, and power factor. One method uses all four variables and the other uses only voltage and current. The test data deals specifically with continuous miners, but the theory could be applied to other mining machines driven by multiple induction motors. Results are given for the tramming, maneuvering, cutting, and cutting with loading operations and are in the form of a minimum, a maximum, and an average of the shaft-torque time function. The theoretical development and computer implementation are discussed for both methods.

Jan 1, 1978

By Vincent R. Bovino

The future of coal mining is bright - only if we realize and then manage as though we are competing in the world marketplace. Accept it or not, our market is global. We tend to think our problems are caused by the uncontrolled foreign production of coal, and the uncontrolled drop in coal prices. I submit that our problem is not the world pricing structure - part of our problem is our on-site cost structure - our productivity. We are the world's best at getting out the tons - that is, production. Most mine managers thrive on that. What is in serious doubt is our productivity - our ability to produce tons of coal in a cost effective manner. Please think about the difference: ? Production is tons of coal. ? Producitivity is decreasing the cost per ton of coal. Our future is bright, only if we can increase our productivity by trying different and new ways to manage our businesses and our people. Unlike the easy successes during the 60's and 70's, success in the decades ahead will depend on our ability to achieve greater levels of productivity. If mining and other old line smoke stack industries are experiencing difficult times in 1986, three years into an economic recovery, what will happen when the economy cycles downward? Any operation can achieve unlimited productivity by developing the resources within their employees. As an industry, we have harvested the

Jan 1, 1986

By John Furness Park

This paper, which describes the mining methods and practices at the Jarbidge mines of the Elkoro Mines Co., is one of a series of papers being issued by the United States Bureau of Mines on mining methods and costs in various districts throughout the United States. LOCATION AND HISTORY The Elkoro Mines Co.'s property is located at Jarbidge in the northeast corner of the State of Nevada (fig. 1). The camp is 9 miles south of the Idaho line; the nearest railroad point is 70 miles to the northeast at Rogerson, Idaho. A road leading to the south and connecting with Deeth, Nev., is passable during the summer months; the heavy snowfall which generally takes place early in the winter usually closes the road until the first of July.

Jan 1, 1931

By L. Holappa

Ferroalloys like FeCr and FeMn are mostly produced in submerged electric arc furnaces. Typical for these processes is that the amount of slag is quite high. The role of slag for the process is thus very essential and specific chemical, physical and electrical properties are required. Slags in FeCr process are Al2O3-MgO-SiO2based with minor contents of CaO and CrOx. Slags in FeMn process are principally similar originating from silicate minerals but can have quite a large composition range depending on the manganese ore and its gangue minerals. Characteristic for FeMn process is that the slag is rich in manganese oxide up to above 30%. A common feature is that the ?ore oxides?, CrOx and MnOxin slag are not only divalent oxides but also containing some higher valent of Cr3+ and Mn3+, respectively. In FeMn production conditions the slag consists mainlyMn2+. This is a clear difference from CrOx containing slags in which both two and three valent coexists in slags in FeCr process conditions. Oxidation state of the slag depends on the ambient oxygen partial pressure and temperature in the examined process. Thermodynamics of slags in ferroalloys production are discussed in terms of phase diagrams, activities and oxidation state of chromium and manganese oxides. Further, slag properties influencing reduction kinetics and furnace operation i.e. viscosity and electrical resistivity are discussed. Keywords: review, slags, ferroalloys, ferrochromium, ferromanganese, thermodynamics, physical properties.

Jan 1, 2004

By Staff

This paper contains an outline of the history of the Sullivan mine, now owned and operated by the Consolidated Mining & Smelting Company of Canada, Ltd.; some account of the various methods by which the ore has been treated in efforts to solve the metallurgical problems which its complex character presents, and, finally, to give, as concisely as possible, an outline of the present mining and metallurgical practice along with a description of the mining and milling plant at Kimberley and the smelter and refineries at Trail, in British Columbia. EARLY HISTORY In May, 1892, Pat Sullivan, John Cleaver and Mike Holland left the Coeur d'Alene country, in the State of Idaho, to seek their fortunes in the Kootenay lake district in British Columbia. They were aided in their venture by James Cronin, afterwards locator and owner of the St. Eugene mine at Moyie. Two months later the party broke up, Sullivan and Cleaver crossing the mountains from Crawford bay on Kootenay lake to the head waters of the St. Mary's river, the course of which they followed down to Fort Steele. Here they found a good deal of excitement about the North Star mine, and, joined by "Ed." Smith, who was familiar with this part of the country, they determined to return and prospect in its vicinity.

Jan 1, 1924

By A. J. Petrina

"At Craigmont Mines Limited, Merritt, B.C., tests were begun in March, 1965 to evaluate the use of shotcrete as a method of ground support m development headings; previously, ail drifts had. required support, which was accomplished mainly by timber and occasionally by the use of rockbolts. The tests demonstrated conclusively the effectiveness of shotcrete for supporting Craigmont's ground. Gradually, the use of timber for ground support was virtually eliminated in favour of shotcrete. This paper describes the equipment used, the operation of the equipment, pertinent cast data and opera ting details. Also discussed are other uses developed for the shot-creting equipment, the use of reinforcing bar rockbolts, and a method used at Craigmont for lining long vertical ore passes with steel plate and concrete as part of the raising cycle."

Jan 1, 1968

By Vincent A. Scovazzo

Many accepted and useful methods have been developed and used for pillar design. Each method has its own particular strengths and weaknesses. Three methods addressed in this paper are; empirical, analytical, and numerical. Each of these methods have their place in pillar design. Analytical methods result in close approximation of the ideal pillar (most economical, stable, or yielding) and are suitable for use in final mine design and in analyzing existing problems. BOYD uses the analytical method to develop site specific mine pillars design equations. An analytical method, one of many developed by BOYD, is presented here and is based on the procedure used by Wilson (1) and others in development of their empirical equations.

Jan 1, 1995

By S. S. Gajin

The paper deals with parametric vibrations of an axially loaded pile. The dynamic model is derived by means of Hamilton's principle and reduced to a partial differen­tial equation with variable coefficients and corresponding initial boundary conditions, taking into account the influence of the axial force. By means of Galerkin's method, the partial equation is further reduced to a Matheu-Hill's matrix partial differential equation of the second order. Since the problem of the dynamic stabili­ty has been expressed only for standing piles, the paper analyses in detail two stan­ding piles, one which is restreined, whereas the other is hinged at the toe. In both cases it is assumed that head horisontal displacement is prevented. Using Ruunge-Kutt's method, a computer analysis in the time domain is given for the characteristic parameters and possible boundary conditions. The paper analyses the stationariness of the process, as well as the dependance of the critical excitation frequency upon the pile parameters. The results have been illustrated by respective diagrams.

Jan 1, 1989

By Walter R. Storms

"INTRODUCTION The New Mexico potash exploration described herein was begun in February 1944 with the object of determining the amount of potash ore in the United States Potash Reserve in Eddy County by exploring the deposit from the surface with vertical drill holes. The work was completed in October of the same year after 16 holes had been drilled on the Reserve and in the immediate vicinity.The Potash Reserve is in secs. 26; 27, 34, and 35, T. 20 S., R. 29 E., Eddy County, N. Mex., about 14 miles northeast of Carlsbad, as shown on figures 1 and 1A, Drilling was done in secs. 26 and 35 of the Reserve, in section 14 to the north, end sections 2 and 56 to the east. All work was done on Governnent-owned, unleased land.The Reserve may be reached from Carlsbad by traveling to the north-east by automobile for about 14 miles on the paved U. S. highway 62. This highway passes through the southern part of the Reserve. Other parts may be reached by a State highway that passes along the eastern side of the area, then by poor dirt roads to various drilling sites. A branch of the Santa Fe Railroad also passes across the southern part of the Reserve enroute from Carlsbad to the plant and mine of the Potash Co. of America, a few miles northeast of the project and 22 miles from Carlsbad."

Aug 1, 1947

By Ned Baker

This report describes the development, construction and testing of a hydrogen-fueled underground mining vehicle. A hydrogen burning engine and a hydride fuel system, both previously developed, were installed on a modified utility vehicle, together with the necessary controls. In tests and in actual use, the vehicle performance in terms of speed, acceleration, and gradeability exceeded that of the standard, unmodified vehicle.

Jan 1, 1985

By M. Board, C. Alexander, M. Ramström, W. Johnson

"Hecla - Lucky Friday is an underground mine that has been in operation for 75 years and produces from depths over 2.3 kilometers (7,500 ft) below the surface. Technical issues at this depth include high rock temperature and seismicity. The use of new innovative mechanization is being examined for ore production. Principal advantages of mechanization are the reduced exposure of personnel, reduced seismic triggers, and increased efficiency. Lucky Friday has partnered with Atlas Copco to design, build, test, and deploy a mechanical miner for primary ore production. We will discuss the background, concept, development and deployment strategy for a mechanical miner. BACKGROUND The Coeur d’Alene mining district in northern Idaho was first discovered over 130 years ago and has been the most prolific silver producer in the US. The Lucky Friday mine of Hecla Mining Company (Figure 1) is the oldest (76 years) and deepest mine in the district, producing approximately 725 tonne per day (800 tons per day) from narrow, vertical lead-zinc-silver veins. Cut and fill stoping is currently employed at depths to about 2.3 km (7,500 ft) below ground surface, making it one of the three deepest mines in the western hemisphere. The #4 Shaft, a winze collared at approximately 1.8 km (5,900 ft) below surface and extending to a depth of nearly 2.9 km (9,600 ft), was recently completed to exploit the high-grade 30 Vein, whose resource has been identified to at least 3 km (9,700 ft) in depth.As with other deep mines worldwide, a number of challenging engineering issues occur, including: high in situ rock stress, squeezing ground, seismicity, high in situ rock temperature that requires refrigeration, diesel equipment that generates diesel particulate matter (DPM) requiring high ventilation airflow rates, and long travel times for worker transport to mining areas."

Jan 1, 2018