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By Michael J. Rinker, R. V. Nicholson

Waste rock and tailings from a proposed nickel-copper mine were evaluated extensively for potential on-land and under water disposal. Standard humidity cell and column tests were conducted over a two year period and underwater tests were conducted for more than one year. Nickel concentrations or release rates from the waste were consistent with nickel to sulphur ratios in the solids. This relationship provided a tool to assess the risk of metal leaching in wastes using nickel and sulphide assays. Nickel release and mobilization was found to occur for both oxidizing leach tests and under water tests even when pH values exceeded 7.0. Geochemical model calculations showed that iron oxyhydroxide solids can scavenge nickel from leachate and that the nickel inventory can be released from solids at relatively lower, but neutral pH values.

Jan 1, 2000

By Karin Bäppler

Mechanized tunnelling is undoubtedly one of the most innovative disciplines in the international construction industry. Technological progress in the planning and production of most modern infrastructure is constantly striving for new records on basis of highly specialized solutions and know-how. Humans and technology have accomplished excellent achievements over the past few years in varied, difficult terrain. New length records are being set with the 64 km long Brenner Base Tunnel and the HS2 Railway system currently under construction in Europe. Cities are building ever longer and more demanding sewer tunnels. The paper highlights some of the challenges, trends and innovations of long-distance tunnelling projects.

Dec 1, 2022

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

The STG integrated method of grouting can be divided into the several phases as outlined in preceding chapters. For convenience in describing technical procedures, they can be grouped as: 1. Investigations in non-geotechnical, exploration bore¬holes. 2. Drilling the holes necessary for geotechnical investiga¬tions and for grouting. 3. Conducting appropriate hydrogeologic tests in those holes and consequent calculations. 4. The preparation of a clay mortar. 5. The preparation of a clay-cement grout with additives. 6. The injection of the grout. 7. Checking the quality of the grout curtain. Items 2, 4, 5 and 6 are discussed below. These activities contain the procedural aspects of grouting. 7.1 DRILLING GROUT HOLES Grout holes belong to a very select class of drillholes. The necessary drilling equipment must enable the operator to drill the holes in inclined, twisting directions. In addi¬tion, the equipment must permit drilling under structurally and hydrogeologically complex conditions. The equipment must permit the operator to conduct specific testing activi¬ties in the drillholes. These activities include hydrodynamic analyses, flowmetric analysis, and the installation and re¬moval of deflectors, packers, grouting plugs, casings, lin¬ers, pumping facilities and other work. As explained in Chapters 5 and 6, the grouting of sat¬urated rock is conducted both through holes drilled from the surface and through holes drilled from the face of a shaft, drift, or tunnel. The drilling of grout holes from the surface can be carried out by an aggregate of equipment that can be the same equipment used for drilling exploration boreholes. Depending on the projected depth of the holes, the STG ZIF- 1200MR and ZIF-650 drilling rigs (or modifications of them) are used. The SKB-4, SKB-5 and SKB-7 high-output drilling rigs (workover rigs) have been used in recent years. The STG BMP-24 drilling rigs are used for carrying out the descent-lifting operations for drilling holes from the sur¬face. In a number of cases, the UKB-500C, URB-3AM and URB-2A power-fed drilling equipment is used for drilling holes from the surface. It should be noted that the drilling of grout holes usually requires an electric drive assembly. Drilling with diesel drive assemblies would be used only to preclude the possi¬bility of losing electric power during drilling operations. Turbine drilling would be used only for appropriate techni¬cal-economic reasons. The selection of the grout hole design is determined by the hydrogeological and structural geological conditions at the site as interpreted in Chapters 2 and 3. The design of grout holes must be as simple as possible in order to min¬imize costs. Hole design guidelines can be adopted as a basis for this purpose, unless site-specific conditions require otherwise. The mouth of the drillhole must be outfitted with a guide-pipe having a length of at least 2 m and an outside diameter from 219 to 234 mm. The upper part of the hole must be attached by a jig consisting of borehole casing that has a diameter ranging from 108 to 146 mm. The length of the jig h is determined by the equation [ ] where k = 1.1 to 1.2 is the load factor; P,, is the injection pressure of the grout into the mouth of the drillhole; D is the external diameter of the jig pipe; m = 0.6 to 0.7 is the work condition factor; T[ ].1 MPa is the bonding value of the rock cemented to the jig. The grout hole is drilled from the jig shoe to the de¬signed depth using a rock-crushing bit with a diameter of 93 mm or more. However in complex hydrogeological condi¬tions when the shaft intersects unstable rock, other hole designs can be used. The diameter of grout holes must permit an aggregate of investigations to be conducted in them using down-hole borehole geophysical and flowmetric logging instruments. It is necessary to cement the casing strings reliably in the grout holes, thereby permitting the injection of the grout under high pressure through the pres¬sure tight mouth of the borehole. Technological details for drilling grout holes such as the rotational rate of the drill bit, the axial load on the rock¬crushing bit, and the drilling fluid flow rate are optimized for the specific rock-hydrogeological conditions largely by experience. The hole diameter and type of rock crushing bit are important variables in these details of drilling. It is advisable to drill holes using water as the drilling fluid. However in those cases where the hole walls are unstable or under very high ground water pressure, the use of a well¬ engineered drilling mud should not be precluded. Unfortu¬nately, drilling mud can influence test results. When steeply dipping fractures with inclination angles of more than 60 to 70° are encountered, it is necessary to drill guided, inclined grout holes. This procedure permits the maximum number of fractures in each hole to be stu-

Jan 1, 1993

By David Wilson, T. V. Rao

INTRODUCTION The Alchem trona mine, owned and operated by Allied Chemical Corp., is located approximately 32 km (20 miles) west of the town of Green River, and 8 km (5 miles) north of Interstate 80 in southwest Wyoming. Trona, a naturally occuring sodium sesquicarbonate (NaCO3 • NaHCO3 • 2H2O), is used for the production of sodium carbonate (Na2CO3), commonly known as soda ash. Present annual US consumption of soda ash is approximately 9.07 Mt (10 million st), out of which approximately 7.3 Mt (8 million st) is produced from the Wyoming trona mines. About 44% of soda ash is consumed by the glass industry; 3 % by exports; and the remaining portion by paper, chemical, soap, textile, and nonferrous industries. Allied Chemical operates an underground mine at an average depth of about 488 m (1600 ft). There are three other trona mines in the area operated by FMC Corp., Stauffer Chemical Corp., and Texasgulf Inc. (see Fig. 1). Allied Chemical became interested in southwest Wyoming trona deposits in the early 1940s. Explora¬tion work was begun in 1959 with the acquisition of federal leases. Exploitation of the reserves began in 1968. Two shafts, one 6.1 m (20 ft) and another 3.7 m (12 ft), were completed prior to 1968. The 6.1-m (20-ft) diam service shaft is divided in half by a cur¬tain wall, with one half used for men and material and the other for ventilation as a return air shaft. During the period from 1968 to 1974, the 3.7-m (12-ft) shaft was used as production shaft with two 7.3-t (8-st) skips. Due to the increased demand for soda ash, plant and mine expansions were completed during 1974. Under this expansion, the material handling system capacity was increased by constructing a new 2721-t (3000-st) capacity ore bin, sinking a 6.1-m (20-ft) diam produc¬tion shaft, and installing a friction hoist with two 22.7-t (25-st) skips. The Alchem mine manpower is 484 employees and about 2.3 Mt (2.5 million st) of trona was produced during 1976. Three different mining systems are used at the Alchem mine: conventional, continuous, and longwall. Allied Chemical introduced its first longwall unit in 1973, which was also the first longwall system introduced in the Wyoming trona district. At present, four conventional, seven continuous, and two longwall units are in operation. This equipment is operated in ten production areas. GEOLOGY The bedded trona and trona-halite in the Green River basin of southwestern Wyoming is a major re¬source for the nation. There are 25 beds that exceed 0.9 m (3 ft) in thickness and underlie areas of about 2590 km2 (1000 sq miles). There are other minor beds that range in thickness from 0.3 to 1.2 m (1 to 4 ft). Allied Chemical's Alchem mine is located in bed 17. The trona bed numbering system starts at the lowest elevation and increases upwards. The trona beds are in the Wilkens Peak member of the Green River formation. The Green River formation was deposited in a lake that began in early Eocene as a large body of fresh water, shrank in size and became saline, and then became fresh water again. The sedi¬ment during the saline phase of the lake, which includes the trona beds, is called the Wilkens Peak member. The underlying freshwater lake deposit is called the Tipton member and includes a 25.9-m (85-ft) sand unit which is a high pressure, low permeability aquifer. The over¬lying sediment is called the Laney shale. The Green River formation is underlaid by the Wasatch formation and overlaid by the Bridger formation (see Fig. 2). The structure of the beds in the area of the Allied Chemical mine is mostly postdepositional asymmetrical folds of small amplitude. These folds range in size of 3 to 6.1 m (10 to 20 ft) to as much as 45.7 m (150 ft). The trend of the folds is northwesterly with the north¬east flank of the anticlines having a steep dip and the southwest having gentle dips. Trona bed 17 varies in thickness from 2.4 to 3.4 m (8 to 11 ft) over the 93.2 km2 (36 sq miles) holdings of Allied Chemical and is flat lying except for the rela¬tively minor folding and a negligible regional dip of about 0.017 rad (1°). Trona occurs naturally in several physical forms in the Green River basin, being both massive and crystal¬line. The trona ore in the Alchem mine is a fine-grained massive ore with a compressive strength of approxi-

Jan 1, 1982

By Hans Ehdwall

INTRODUCTION Investigations of radon and radon daughter concentrations in Swedish [non-uranium] mines started in the late 1960's. The first screening measurements showed that the average annual exposure to radon and radon daughter products was 4.7 WLM. The main reason for high radon and radon daughter concentrations was inefficient ventilation and radonrich water entering the mine. In the radon regulations worked out later it was stated that no miner should be exposed to more than 60 000 pCi h/1 equilibrium equivalent concentration of radon annual exposure, corresponding to 3.6 WLM. Now, 1981 the situation has changed considerably. From the average annual exposure of 4.7 WLM in 1970 it is now only 0.7 WLM. Sweden has up to now had only one [uranium] mine and the work there has only been investigative. However, there are plans for a commercial uranium mine in another part of Sweden. The radon problems in these mines are widely different depending on the mineralogy. NON-URANIUM MINES The radiation problems in Swedish mines were not recognised until the late 60's. The first radon and radon daughter measurements were made in some sulphide ore mines in 1967 (1). The radon and radon daughter concentrations were surprisingly high for non-uranium mines. In order to have a complete picture of the radon situation in Swedish mines the National Institute of Radiation Protection (NIRP) decided to make measurements in all, at that time about 60 mines (2). To get results as fast as possible measurements on radon gas seemed most appropriate to start with. Sampling was made by mailing a number of evacuated 4.8 litre conventional propane containers from NIRP to each mine. The containers were then opened at the place of interest. After sampling the containers were sealed and then mailed back to the institute for measurement. The measurements were made in ionization chambers. This method only gave the radon concentration and the radon daughter concentration was estimated by multiplying the radon concentration by an assumed equilibrium factor. The equilibrium factor is defined as the ratio of the total potential alpha energy for the given daughter concentration to the total potential alpha energy of the daughters if they are in equilibrium with the given radon concentration. The results of this first preliminary survey indicated that a great many of the Swedish miners probably had an annual radon daughter exposure of more than 3.6 WLM. As the radiation exposure in non-uranium mines was not regulated in either the Swedish Radiation Protection Act or the Swedish Labour Protection Act work was started on special radon regulations. A lung cancer mortality study was also started. To check the results of the first survey and to get experience and knowledge of radon problems in mines, it was decided that personnel from the NIRP should visit each mine for a detailed investigation of radon and radon daughter concentrations starting with the ones with the highest radon concentrations. The main reasons for these so-called "basic measurements" were: 1. To estimate the doses received by Swedish miners 2. To find the sources of the high radon and radon daughter concentrations 3. To find appropriate counter-measures 4. To determine the most typical equilibrium factor for each mine. Unlike most uranium mines the reason for high radon concentrations in non-uranium mines is seldom the occurrence of highly radioactive minerals. The main sources were found to be waste-rock and radon-rich water. In order to filter and warm up the inlet air, especially in winter time, it was very common at that time to suck the air through broken wasterock. By doing so the air was contaminated with radon from the waste-rock and radon-rich water in it. It is noteworthy that the radium and uranium concentration in the waste-rock is relatively low. The uranium concentration is only of the order of 15 - 20 ppm. The action to prevent this contamination of the inlet air was to change the direction of the ventilation and in the case of radon-rich water entering the mine the action was to prevent the air coming into contact with the water. The first calculation of the radon daughter exposure of Swedish miners was based on radon gas measurements. The radon daughter concentration was estimated by using an assumed equilibrium factor of 0.5. Later when the mines were visited by institute staff it was possible to compare the assumed equilibrium factor with the measured ones. It was found that the factor varied from 0.15 at the air inlet to 1.0 at the air outlet and the average equilibrium factor on workplaces for almost all mines was between 0.4 and 0.6. The result of the exposure calculation in 1970 showed that more than 40 % of the miners had an annual radon daughter exposure of more than 3.6 WLM. The overall average was 4.7 WLM and the maximum annual expo-

Jan 1, 1981

By Jack V. Hill

A unique type of grinding mill liner has been developed by Waagner-Biro A.G. of Austria. This liner system, termed an angular spiral lining (.ESL), was, successfully tested by Magma Copper Company's San Manuel concentrator in both their 10 Ft x 10 Ft (3.05m x 3.05m) and 12 Ft, 6 In x 14 Ft (3.81m x 4.27m) grate discharge wet grinding ball mills. Magma Copper's savings in power and grinding balls justified conversion of more than half of their twenty-six ball mills to the ESL system.1,2 The ASL system was provided by the Capitol Castings Division of Midland Ross Corporation, the exclusive licensee for such conversions. Capitol Castings proposed the following as expected benefits from installation of an ASL system in an 18 Ft x 21 Ft (5.49m x 6.40m) overflow discharge wet grinding ball mill at the Pinto Valley Operation of Cities Service Company: "(a) Reduction of power consumption per ton of ore milled = 12%, (b) Increased throughput = 6%, (c) Reduction of ball consumption = 6%."

Jan 1, 1982

By B. L. Gibbs

The mining industry is constantly in a state of flux -boom or bust; mine openings and closings. Part of the reason is the nature of ore deposits -they are there or they are not. Once they are mined, they are reclaimed and the move is on to the next deposit. Risks are usually high, and payoffs can be great or not so great. Many factors go into making a property payoff, and mining companies are continually searching for ways to remain profitable as long as possible. Computers and the people who can make them work are one of the ways available to improve productivity and ensure this happens. Computers have been used in the mining industry for the last 30 years (and more in a few cases). During the past 10 years, the explosion of micro-computing has changed how many mining companies do their work.

Jan 1, 1993

By Panagiotis Chryssanthakis

"The aim of this study is to give a better understanding for the performance of the composite material RRS and the synergy effect of its components under different stress regimes and while an excavation face advances during tunneling. The lower part of the Q-system scale (i.e. Q-values of 0.4 or less for a tunnel span of 10m) is commonly used for tunnel applications in softer rock formations. This is implied in the updated Q-chart, where the dimensioning of rock support by using Reinforced Ribs of Shotcrete (RRS) comprises the use of fiber reinforced shotcrete, S(fr), rock bolts and shotcreted beams reinforced with steel rebars. In this study, an effort has been made to simulate RRS in a numerical code and to study its behaviour in a real case study. A double tube, triple lane tunnel in weathered tuff has been chosen as reference material for testing RRS under changing load. A rather extensive parametric study has been run by using the same key input rock mass parameters and three different in situ stress regimes (sH/sv =2, sH/sv =1 and sH/sv =0.5). Changes have only been made to the RRS key input parameters in order to simulate the changing loads during the face excavation process in tunneling. The critical part of this numerical modelling work that is under investigation in this study under each stress regime, is the period just after the excavation: i.e 4 hours; 16 hours after excavation; and finally after the complete curing (28 days) of the cement based components. The RRS are installed in three consecutive numerical steps as follows: a) the application of S(fr) on the rock surface in order to even it up for the RRS application; b) radial bolting; and c) installation of the reinforcing steel rebars, that form the reinforced beam. The results are varying and are strongly dependent on the imposed stress regime and they show in general the following tendencies: The axial forces on S(fr) show that S(fr) is almost overloaded during the first 4 hours in all 3 stages. More particularly: For the case with sH/sv =2 a) the axial forces on S(fr) and on the steel rebars are reduced after 16 hours and 28 days; b) the same is observed on the shear stresses in S(fr) and on steel rebars, they are decreasing after 16 hours and 28 days; and c) the bolts loads are almost unchanged. For the case with sH/sv =1 a) the axial forces on the S(fr) are reduced after 16 hours and 28 days; b) the same is observed for the axial and shear stresses, as well as moments on the steel rebars, they are decreasing after 16 hours and 28 days; and c) the bolts loads are almost unchanged."

Jan 1, 2016

By Jim McKinley

With a growing number of mines reaching closure and seeing closure costs materialize, regulators, lenders, and banks are observing projects that fail economically by miscalculating these costs. Mine closure costs are often categorized as a type of liability which is defined as the money owed by the company to offset effects from the project, and the money available to pay for those debts (e.g., bond). Often, decisions made before Day 1 of mining (e.g., co-disposal, dry stack tailings, waste segregation) can significantly change the bond calculation. Once the bond is calculated and funded, operators are encouraged to conduct closure research and use applied science for progressive bond release prior to final closure. However, this author has observed few instances of financially significant progressive bond release, as there are typically no robust procedures for bond return, either before or during closure activities. As a result, many companies decide that entering a state of Care & Maintenance (C&M) is less risky than attempting to fund closure activities with an unknown bond return mechanism. Although perhaps less risky in the short-term, the primary long-term risk with entering C&M is that it increases the chances of the bond not covering the cost of closure and the chances of the government/regulatory agency inheriting the mine closure. Bonds are typically reevaluated on a regular cycle (e.g., five years) so every mining company goes through this financing challenge eventually.

Feb 22, 2026

By Jan Wolf

The iron ore industry presently utilizes two types of de-posits. The first are the high grade deposits generally located in remote areas and, secondly, the low grade de-posits which are normally found nearer to the steel-making centers. The high grade deposits produce a direct shipping ore which has an iron content which may be close to the maxi-mum 70% of Fe. These deposits have reserves which are ex-pressed in billions of tons. The ore needs little or no up-grading and where this is performed is restricted to ore washing. With the large demand for iron ore pellets, plus a national drive to perform as much of the upgrading as possible in the native country, there is a trend to erecting pellet plants at the mine location~ Large iron ore suppliers such as Australia, Brazil, Venezuela and Liberia already have pellet plants in operation, and there are new plants projected or expansion projects under study. The low grade ores are mined at locations closer to steel mills such as in Minnesota, Michigan and Canada. These low grade ores require methods of beneficiation to obtain a commercially accept-able product with the modern beneficiation techniques it is no~, possible to produce iron ore pellets of 68% Fe and 4.5% silica from ores containing as low as 26% Fe. Figure 1 gives a general picture of the types of iron ores and their beneficiation. The final product of the high grade ore is a coarse concentrate, sinter product or pellets. The final product of the low grade ores is pellets and, in some cases, coarse concentrates.

Jan 1, 1971

By Dave Southworth

After Abe Lee discovered gold (in April 1860) near the site of present-day Leadville, Colorado, he and is group staked nearly all of California Gulch with "speculative claims." They did so in an effort to grab any property that might be potentially profitable. But the plan did not work. New prospectors demanded a miners' meeting, from which arose the official "Bylaws of California Mining District, California Gulch, Arkansas River." These bylaws, adopted on May 12, 1860, established provisions as to the number, size and type of claims that could be filed. Fair laws invited more newcomers. By mid-summer, 5,000 people with tents and wagons were scattered from one end of the gulch to the other.Among the earliest arrivals to the region that spring were Horace Austin Warner Tabor and his wife, Augusta. H.A.W. Tabor, who was to become a pillar in the development of Leadville, arrived at California Gulch during a food shortage. He sacrificed his oxen so that hungry miners could eat. By doing so, he immediately became the friend of many. Augusta established a little store where she sold her baked goods and provided meals. Like most of the men, Horace prospected the surrounding hills. He only did so, however, for a short time. California Gulch yielded more than $5 million in gold during the next five years. By 1866, however, the placer gold was gone and most of the miners had packed up and left. Charles J. Mullen and Cooper Smith, who were grubstaked by a Philadelphian named J. Marshall Paul, found a new, rich lode at the Printer Boy Mine in 1868. It was the first underground gold mine in the Leadville area and marked the transition from surface mining to underground hard rock mining. Its discovery created a new but modest flurry of activity. Heavy black sand covered the region and made gold mining difficult. In 1875 William Stevens and Alvinus Wood decided to have the sand assayed. They discovered that it contained 15 ounces per ton of silver and was rich in carbonate. Stevens and Wood kept their find a secret for nearly two years. When the word spread in late 1877, the silver boom was on. In 1878 prospectors, merchants and others trickled into town from all directions. Dr. David H. Dougan closed his office. in Alma, crossed Mosquito Pass and set up shop in Leadville. Nobody seemed to need a doctor, however, and Dougan sat in his office for 28 days without a customer. On the 29th day he received word that there had been a mining accident and he was needed immediately. -That was the begin¬ning of his illustrious career in Leadville. Dougan was a successful physician, became mayor in 1881 and was later president of the Carbonate National Bank. Leadville was a simple community of log structures in 1878. By June of the following year, however, it had blossomed considerably. According to Cass Carpenter, as of May 1,1879, Leadville had "...19 hotels, 41 lodging houses, 82 drinking saloons, 38 restaurants, 13 wholesale liquor houses, 10 lumber yards, 7 smelting and reduction works, 2 sampling works for testing ores, 12 blacksmith shops, 6 livery stables, 6 jewelry stores, 3 undertakers and 21 gambling houses where all sorts of games are played as openly as the Sunday School sermon is conducted." Additionally, there were 36 brothels. When George E. King and other architects arrived, Leadville took on an air of elegance and sophistication. In sharp contrast to some of the Victorian-style, gingerbread-laced homes in Leadville, its suburbs were comprised of many tenement shacks and plain cabins mixed with a few commercial buildings, many mines and smelters. Mines sprinkled the area east of Leadville like pepper from a pepper shaker. Wherever there was a cluster of mines, a camp usually sprang up. Most were very close together. South of Leadville the giant Arkansas Valley lead smelter sprawled out between Stringtown and Bucktown in an area known as Smelter Valley. Founded in 1879 as the Billing and Eilers Smelter, the Arkansas Valley plant was operated by the American Smelting and Refining Company. Several other smelters operated in the vicin¬ity. In April 1878, Horace Tabor grubstaked prospectors August Rische and George Hook. For $64 worth of supplies and tools, he was guaranteed a one-third interest in their findings. Rische and Hook discovered the Little Pittsburg (often spelled Pittsburgh) Mine and made Tabor an instant millionaire. Tabor once purchased a "salted" mine, the Chrysolite; but instead of considering it a bad deal, he further developed the property and struck rich silver ore. Later, Tabor purchased the Matchless Mine for $117,000, possibly the only investment he made without partners. During its peak years, the Matchless yielded $1 million worth

Jan 1, 1998

By C. E. Semler

A multiyear program involving online, in-vessel exposures at Barrick Goldstrike Mines Inc. (BGMI) has provided information and experience about the materials and designs for brickwork linings in pressure oxidation autoclaves. A variety of brick and mortar products was evaluated. Properties before and after exposure were used to document the performance. Some benefits include the development of a new silicate mortar, a better understanding of brick requirements and increased autoclave operating time between maintenance shutdowns. BGMI operates six pressure oxidation autoclaves at its Eureka County, NV gold operation. These autoclave process 15.8 k/ld (17,500 stpd) of gold ore. Figure 1 is a flowsheet of the six autoclave circuits. The gold ore is refractory, in which the gold particles are mineralogically encapsulated on a micro- and submicroscopic level. The autoclaves are 4.4- m- (14-ft-) diam and 26 m (85 ft) long (tip-to-tip). Inside of each is a full brickwork (brick plus mortar) lining that divides the vessel into five compartments. Each compartment has a motor-driven agitator that keeps the slurry suspended and disperses the oxygen to enhance reaction. During autoclave operation, the conditions are steady state at 210° to 232° C (410° to 450° F) and 61 to 62 kPa (420-430 psi). As shown in Fig. 1, there are other accessory vessels (splash and flash towers) in the autoclave circuits that are also brickwork-lined. The linings used to date, other than small test sections, have all been acid-resistant fireclay bricks, bonded with alkali silicate mortar, PbO- based mortar and/or vinyl ester mortar, but mainly alkali silicate mortar.

Jan 1, 1997

By E. A. Barrett, L. L. Rethi

Throughout all segments of the United States mining industry, 163 independent contractor employees were fatally injured between 1990 and 1997. Factors that contributed to these injuries were identified after reviewing descriptive information contained in Mine Safety and Health Administration fatality reports for each of the incidents. Thirty-three unique contributing factors were identified and multiple factors were noted for each fatality. Those most frequently discerned were "failure to follow `standard operating procedures"', (35% of the fatalities), "poor positioning of employee" (26% of the fatalities) and "failure to maintain equipment" (23% of the fatalities). Contributing factors were then associated with the accident and job classifications of the victims. Factors having the highest frequencies in the top five accident classifications where most fatalities occurred were as follows: powered haulage - "failure to maintain equipment"; machinery - "poor positioning of employee"; slip/fall of person - "inadequate fall protection" and "failure to wear safety line"; electrical - "equipment contact with power line"; and falling/rolling/sliding rock - "poor positioning of employee. " Among the job classifications in which most fatalities occurred, leading contributing factors were as follows: truck driver - "failure to follow standard operating procedures" and "poor positioning of employee"; laborer/utility man/pumper - "poor positioning of employee"; mechanic/ repairman and welder - "failure to follow standard operating procedures." Recommendations for utilizing contributing factor information as a basis for safety training are proposed. Training programs may be developed that inform independent contractor employees of factors which, in the past, have contributed to fatal injuries in their job classifications. Then, by combining this information with contributing factors identified in "related" accident classifications, a truly focused training intervention that addresses very specific workplace hazards may he developed.

Jan 1, 2002

By R. Brändle, R. Luis-Fonseca, R. Cornejo, G. Fischer, G. Von Rickenbach, J. Jarufe

Rock damage around underground excavations has been an important hazard in mines and tunnels. Rock dilation generates buckling that needs to be contained before it causes damage to equipment or safety concerns to mine workers. In seismically active mines, rockburst occurrence has been widely recognized as one of the most important hazards in deep mines and tunnels. While there is a wide range of cases where rockburst hazard can be mitigated through dedicated excavation sequences, reentry procedures or exclusion zones, there are still many situations where large unexpected rockbursting may occur, relying only on rock support as the last mitigation measure to prevent production and safety losses. Considering the previous situations, surface support as mesh, plays a major role containing the failed/ejected rock and providing a safe working environment to the underground operations. While there have been considerable achievements in improving mesh resistance, with more than 26Kj/m2 of energy dissipation capacity, there is still a lack of research in the overlapping mechanism and load transfer between contiguous mesh elements. This paper presents the state of the art in static and dynamic analysis of welded and rolled mesh overlaps, based on experimentation carried on by Geobrugg in their Research Center, leaving clearly established that the difference in meshes is not only in levels of energy absorption, but also concluding that in the case of high resistance rolled meshes overlapping is not necessary and the transmission of stresses can be fully guaranteed by using suitable connecting elements. All these characteristics, supported by operational advantages as light weight and mechanized mesh installation, provides a considerable improvement in safety and productivity, improving the economic value of the project

Apr 26, 2022

By James R. Yancik

Economical and environmental pressures necessitates the recovery of the minus 0.6 mm x 0 coal generated by todays mining machines. This size fraction, dependent on the type of production unit, maintenance of the production unit, and the Hardgrove Grindability Index, can be a substantial portion of the Preparation Plant feed. Historically, dewatering of the collected 0.6 mm x 0 product has often been rather expensive and marginal with respect to surface moisture reduction. The Freeman United Crown II Preparation Plant more than adequately demonstrates that the continuous scroll type centrifuge can dewater 0.6 mm x 0 coal to acceptable surface moisture levels.

Jan 1, 1981

By James M. Link

The first rule of mineral deposits always seems to be: the deposit is never close to the market. With gems and precious metals, this is not much of a problem. But with most of the mineral commodities used in today's world, transportation becomes a major part of the cost to the consumer. In Japan, for example, more than 60% of the cost of coal to the consumer may be attributed to transportation. The purpose of this article is to examine pipeline transportation, particularly mineral slurry pipelines, as a means of getting the mineral to market. Pipelines Today Regulated pipelines in the US totaled nearly 724 Mm (450,000 miles) in 1980 according to the Federal Energy Regulatory Commission (FERC). Worldwide, more than 210 Mm (125,000 miles) of new pipelines will start construction in 1982 at an estimated cost of more than $150 billion. According to the Bechtel Corp., the period between 1982 and 2000 will see major growth in pipelines of all kinds. In North America, major movements of crude oil from Alaska and synfuels from the Rocky Mountains will require the construction of nearly 16 Mm (10,000 miles) of new pipelines. Nearly 42 Mm (26,000 miles) of new gas transmission pipelines will be needed to tap both US and Canadian arctic gas fields and move the gas to major markets on both coasts. Other pipelines will be required to move additional quantities of natural gas from the Rocky Mountains to more populous regions. Bechtel says more than 24 Mm (15,000 miles) of slurry pipelines will be needed to transport coal from both eastern and western coal provinces to the Mississippi valley and coastal areas. In 1980, 125 companies delivered more than 1 km3 (6.5 billion bbl) of crude oil and 652 hm3 (4.1 billion bl) of products in the US. At the same time natural gas pipelines transported 498 km3 (17.6 trillion cu ft). The reported investment by these companies was nearly $20 billion at the end of 1980. Advantages of Pipeline Transportation The obvious success and vitality of the oil and gas pipeline industry is based, at least in part, on the fact that pipelines are a very efficient and low-cost method of transportation. This fact, coupled with the need for lower cost trans¬portation, has led to the marriage of oil and gas transmission technology and the bulk mineral solids transportation industry. The off-spring of this marriage is the mineral slurry pipeline. Slurry pipelines have a number of advantages over other transportation methods. One of these is the fact that they are buried-out-of-sight, out-of-mind. Second, they are relatively small users of labor because they lend themselves to automation and remote, or even computer control. Third, they offer an attractive economic alternative to other transportation systems. For example, for a 1000 km (621 miles) distance, rail costs of 1?/ km (1.7?/ton-mile) are about equivalent to slurry pipeline costs. But, as the distance increases, pipeline cost per t-km continues to drop while equivalent rail charges remain essentially insensitive to distance. Where existing rail, barge, or ocean ships are available, the cost of new construction associated with a slurry pipeline probably will render it noncompetitive. However, where new rail or other construction is needed, the cost of a slurry pipeline is very competitive. About three-fourths of the cost of a pipeline is in pipe, fittings, and construction. Nearly a fifth of the investment is in pump stations and the remainder is in right-of-way, surface facilities, utility acquisition, communication facilities, and other support areas. The fact that pipelines are capital intensive is, at the present time, a mixed blessing. For example, the delivered cost of coal in a hypothetical project doubles when the cost of money rises from 9% to 17.5% per year. The largest cost element for delivered coal in this example is in depreciation and finance charges. The next largest cost element is electricity to power the pumps, with labor making up the smallest increment of the delivered cost of coal. World Slurry Pipelines The idea of a slurry pipeline was probably first investigated in the latter 19th century, but the first ones were successfully built and operated in the US in the early 1900s. Today, they are fairly common worldwide. Table 1 shows a number of the world's slurry pipelines. An examination of these will further emphasize the fact that pipelines provide a cost competitive alternative where new construction of a transportation system is required. In each of the examples, the desired mineral commodity is located in a remote corner of the world, markets are at a great distance, value of the commodity is not great, and no other transport system is available.

Jan 10, 1982

By J. H. LaVerne Palmer, W. (Bill) C. Cox, Lorant B. Geller

An erector transporter (ET) has been designed and built to lift, carry and erect heavy machinery components, in particular the components of a 7.6 m diameter tunnel boring machine (TBM), within the access tunnels of a submarine coal mine in Eastern Canada. The ET is capable of manoeuvring within a 7.6 m tunnel while precisely positioning components as large as 6 m in diameter and weighing up to 27 tonnes each. The unit can negotiate 20% grades while carrying loads of 30 tonnes. The major design features, verification tests and performance trial results are presented. The versatility of the ET is such that many other applications are possible.

Jan 1, 1985

By Ross Goodman, Jesse Salai

The Ohio River Tunnel Project encountered multiple unexpected geotechnical challenges such as crown and invert instability due to the overstress condition affecting the strata above, below, and in the tunnel alignment during and after TBM excavation along with the potential for methane inflow. This paper showcases how the orientation of a tunnel alignment has a significant impact on how in-situ geological stress can impact the excavation of a tunnel from the contractor’s point of view and how geological features influence gas inflow to the alignment over time.

Jun 13, 2021

By Daniel R. Stewart

The function of the geological engineer is to assist in mine planning and operations by collecting, interpreting, and applying geologic data to the solution of engineering problems. The geologist supports this effort by collecting additional geologic information specifically useful to the engineer. This use of geologic information reinforces the value of mine geology to the mining process beyond the traditional role of orebody definition. Work discussed includes: evaluation of caving strategies at the Climax and Henderson mines; support definition and slope stability at the Climax mine; and some uses of geology in cavability estimation for the Mount Emmons project.

Jan 1, 1983

By Nathaniel Wollman

About nine years ago the President's Materials Policy Co commission under the chairmanship of William Paley released its report entitled "Resources for Freedom." Not long ago the Senate Select Committee on National Water Resources under the chairmanship of Senator Kerr released a study entitled Water Supply and Demand. Both reports contained projections of future water user the report for 1975, and the Select Committee print for 1980 and 2000, It is my task to compare the two reports. In the initial letter from Mr. Kazmann arranging for thus session he suggested that I present "a critical evaluation of the Paley Report's section on water ?ten years later,? which might include a comparison of predicted and actual water-use rates and would spotlight the assumptions which lead to errors in the predictions , . ." It is possible to compare the results of both studies, and it possible to explain art of the differences that can be found between them. The remainder must remain shrouded in doubt, since the Paley Report did not give all details of the methods used in estimating the com¬ponents of water use.

Jan 1, 1961