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By Glen M. Bellew, James F. Mehnert

"The Kansas City District-United States Army Corps of Engineers has completed the design modification of a pile supported floodwall located along the Missouri River in Kansas City, Kansas. This deep foundation supported floodwall protects over 2 billion dollars in infrastructure, is approximately 17 ft (5.2 m) tall and is 1,446 ft (440.7 m) long. Design modifications included the addition of a row of drilled shafts and the addition of relief wells. Typically, deep foundation design is controlled by total stress analysis and undrained soil strength. Neither controlled design at this site, where effective stress and drained strength controlled. During high river stages, artesian pressures develop, leading to a reduction in both effective stress and pile capacity. Deep foundation analysis software available to the designer does not allow these effective stress conditions to be directly entered, so a unique way was developed to account for them. A load test was performed to verify the design. During load testing, the supporting ground was unsaturated and had effective stress conditions different from those that controlled the design. A load test evaluation process involving unsaturated soil mechanics was developed and performed. This paper will share both the effective stress design process used and the load test interpretation process performed.INTRODUCTIONThe Fairfax-BPU Floodwall is part of the Fairfax-Jersey Creek Flood Protection Unit in Kansas City, Kansas. The floodwall was built in the 1940s and currently provides Missouri River flood protection for over 2 billion dollars of infrastructure in the Fairfax Industrial Area. The floodwall averages 17 ft (5.2 m) in height, is 1,446 ft (440.7 m) in length and is divided into 35 separate monoliths. The originally constructed foundation support for each of these monoliths typically consisted of three rows of seven tapered steel monotube piles, each roughly 20 ft (6.1 m) in length. All monontube piles are tipped in sand, well above bedrock."

Jan 1, 2016

By Takashi Sasaoka, Pinyo Meechumna, Hiroshi Takamoto, Pipat Laowattanabandit, Akihiro Hamanaka, Kikuo Matsui, Hideki Shimada

"The EGAT Mae Moh coal mine is an open-cut coal mine in Thailand that produces about 16 million tons of lignite annually to generate 2,400 mw of electricity. The surface mine pit covers is 4 by 7 km (2 by 4 mi) with mining depths ranging up to 260 m (853 ft). A large pit slope is formed with the progression of the mining operation. However, massive slides in the pit slope often occur due to weak planes, such as faults and bedding, and the weak mechanical properties of the rocks. Hence, a 200-300 m (656-984 ft) boundaries of rock block including coal seams is left in front of faults in order to prevent slides and maintain the stability of the pit slope. As a result, there are significant coal reserves beneath the abandoned area along the pit slope.The focus of this paper is on the applicability and design of the highwall mining for recovering much of the coal left along the pit slope.INTRODUCTIONThe EGAT Mae Moh Mine is located about 630 km (186 mi) north of Bangkok, Thailand (see Figure 1). It is the largest coal mine in Thailand and produces about 16 million tons of lignite annually from open-cut mining (70 % of the total coal produced in Thailand). The coal is used to generate 2,400 mw of electricity at the Mae Moh power plant, which is situated near the mining area. The mine pit covers an area of 4 by 7 km (2 by 4 mi) at varied depths up to 260 m (583 ft) as shown in Figure 2. As the mining operation progresses, the pit becomes to be deeper and deeper. In the year 2035, the bottom of the pit is expected to reach depths of 500-600 m and the open-cut mining operation will be finished. The possibility of an underground mining operation is being investigated.A large pit slope is formed with the progression of open-cut mining. However, massive slides often occur in the pit slope along weak planes including faults and bedding planes due to the weak nature of the rock. As a result, 200 to 300 m (656 to 984 ft) boundaries of rock including coal seams are left in place in front of faults to prevent slides and to maintain the stability of pit slope. Moreover, the dip of the slope is less than that of the planned design. As a result, there significant coal reserves abandoned along the pit slope."

Jan 1, 2010

By Maria I. De Rosa

Table 55 and figure 19 show the number of fires and fire injuries for stone mills by state during 1990-2001. Table 55 also shows the injury risk rates, employees' working hours, and lost workdays. In all, 118 fires occurred in 33 states. Seventy-six of the fires caused 82 injuries (including 8 fires and 6 injuries involving contractors). The yearly average was 9.8 fires and 6.8 injuries. The Ewhr value was 873 x 106 hr (In = 0.019), and the LWD value was 1,911. Missouri had the most fires (16 fires and 9 injuries), followed by Pennsylvania (15 fires and 14 injuries), Michigan (9 fires and 5 injuries), and Virginia (7 fires and 5 injuries). Of these states, Virginia had the highest injury risk rate value (In = 0.09 1). Table 56, partly illustrated in figure 20, shows the number of fires, fire injuries, risk rates, employees' working hours, and lost workdays by time period. The number of fires and fire injuries decreased during most of the periods, with a sharp increase during the last period. The number of employees' working hours increased throughout the periods. The In values follow patterns similar to those shown by the injury values. Tables 57-63 show the number of fires by ignition source, method of detection and suppression, equipment involved, location, and burning material by time period. Figure 21 shows the major variables related to fires for 1990-2001. Table 63 shows the fire injuries per number of fires causing injuries and total fires by year, ignition source, equipment involved, and location.

Jan 1, 2004

The meeting of the Board of Directors, on November 22, was attended by eight directors, the Secretary of the Institute, and eight guests. The proposed James Douglas Prize and Tablet was referred to a committee of eleven. Robert H. Richards was appointed representative of this Institute to attend the meeting of the Society for the Promotion of Engineering Education. J. V. N. Dorr was made chairman of the Finance Committee in place of George D. Barron, resigned. The sum of $200 was appropriated for the 1918 expense of the American Engineering Standards Committee; the question of an appropriation for 1919 was referred to the Executive Committee for study and report. The sum of $50 was appropriated to the Treasurer's clerk. The following representatives of the Institute were elected: W. L. Saunders to succeed himself as United Engineering Society Trustee; R. M. Raymond to succeed himself as member of Engineering Foundation Board; S. J. Jennings to succeed himself as member of Engineering Council; George F. Kunz to succeed himself as member of Engineering Council; C. F. Rand to succeed himself as member of John Fritz Medal Board of Award; E. Gybbon Spilsbury to succeed himself as member of the Library Board. To members of the Canadian Mining Institute residing in Canada or Newfoundland, the privilege was extended to subscribe to the Bulletin at the price of $5. An invitation was extended to the Mining and Metallurgical Society of America to consolidate with the American Institute of Mining Engineers. The appointment of Engineering Societies' Employment Bureau was reported. Thirty-nine members and ten associates were elected. Two members were reinstated. Extension of time for the payment of clues was granted to four members. Seventeen members on active duty at the front had their dues remitted.

Jan 1, 1919

The uphole bench extraction strategy at EKJV Management Pty Ltd?s Raleigh Mine, requires simultaneous horizontal and vertical exposure of paste filled stopes. In order to prevent excessive ore dilution, the paste fill must be of sufficient strength to prevent failure during exposure. A numerical modelling methodology has been developed to accurately simulate the initial stress distribution within paste filled stopes, together with the strain-softening loading response of paste fill material. Due to the narrow orebody width and high deformation behaviour of the host rock mass, high closure strains are likely to cause a stiff, brittle fill material to undergo crushing failure upon exposure. The stability of paste fill exposures at the Raleigh Mine was determined by conducting a series of three-dimensional numerical models that incorporate the extraction, filling and exposure sequence of paste filled stopes. Stability charts for 2.5 m and 3.5 m wide simultaneous horizontal and vertical exposures have been derived based upon the numerical modelling results. Implementation of the stability charts has proved successful in limiting the amount of ore dilution from paste fill and provides a basis to optimize paste fill cementation and stability throughout the life of the mine.

May 1, 2007

Jan 1, 1971

By R M. Stephenson

Large diameter raise boring is a common raise excavation method in the Australian mining industry. Geotechnical investigations were conducted on several large diameter raise bore projects at Gwalia Mine to assess the potential for instability and potential back-reaming problems. A thorough geotechnical analysis and subsequent back-analysis was conducted on a vent raise excavated at 5.5 m diameter and 810 m length.The geotechnical investigation for the vent raise included collection of data from an acoustic televiewer survey that was conducted on the pilot hole. The survey results provided structural data and an insight into the effects of stress on the pilot hole. Structural analysis was conducted using a combination of information obtained from the acoustic televiewer survey and geotechnical logs of the diamond drill core drilled close to the site.Further information was collected from drillers’ plods from pilot hole drilling and back-reaming to conduct a back-analysis of the completed shaft, and this was compared with the structural analysis.The key finding from this review was that the acoustic televiewer survey of the pilot hole accurately identified locations with blocky conditions. This is an important supplement to the information available from the main geotechnical investigation hole, as this was offset from the shaft location over most of its path.CITATION:Stephenson, R M and Sandy, M P, 2014. Structural analysis of acoustic televiewer logs for large diameter raise bored shafts, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 89–96 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By J. A. McLennan, J. M. Ortiz

Nickel laterite deposits are typically formed from tropically weathered mafic-to-ultramafic complexes. The resulting nickel concentration is found within soil horizons and is mineable with regular dozer units. At the time of mining, digging elevations are needed at selective mining units (SMUs) or representative dozer regions (RDRs). Moreover, for mine planning, the costs due to dilution and lost ore require mining elevations to be specified. A conditional simulation approach is used to calculate “optimum” RDR mining elevations. Multiple realizations of the ore base surface are simulated. For a single bottom surface realization, dilution and lost ore costs are calculated over the entire vertical extent of each RDR. The optimum mining elevation is the elevation where the lowest summed cost of “accepting dilution” and “wasting ore” occurs. The reported RDR dilution is the cost of mining the volume of waste located below the bottom surface and above the optimum mining elevation. The reported RDR lost ore is the cost of not mining the volume of ore above the bottom surface and below the optimum mining elevation. For all ore bottom surface realizations, optimum RDR mining elevations and their associated dilution and lost ore costs are calculated and compiled to represent production uncertainty.

Jan 1, 2006

By J. A. MacLellan

"CONSTRUCI'ION OF SHERRITT'S CONCENTRATORSSherritt Gordon is operating two mills north of the 56th parallel, and is presently engaged in building a third. The author was closely involved in the design and construction of all three of these concentrators.The Lynn mill is the oldest of these three concentrators-it is a 4,000 ton per day nickel-copper plant built in 1953. The Fox mill is a 3,000 ton per day copper-zinc concentrator built in 1970. The Ruttan mill will be a 10,000 ton per day copper-zinc plant scheduled to go on line in July 1973.LYNN MILL The Lynn mill including the crushing, concentrating, drying and load out plant is about 69,700 square feet or 17.4 square feet per ton per day. '!his plant was originally built in 1927 at Sherridon, 150 miles south of Lynn Lake, as a 2,000 ton per day copper-zinc plant. In 1952 the plant was moved by tractor train to Lynn Lake. The whole plant was re-oriented within the original perimeter walls as a 2,000 ton per day nickel-copper plant. This plant was expanded to 4,000 tons per day by the addition of t1-Jo rod mills and a 5-1/2 short head crusher in 1959.The Sherridon mill processed its last ore in Sherridon in September 1951 and the first load of nickel concentrate was loaded at Lynn Lake on November 9, 1953 when the railroad arrived at Lynn Lake.The site for the Lynn Lake mill was cleared of overburden in the summer of 1952 and foundations for structural steel and SO!l8 equipment foundations were poured the late summer and fall of 1952 • Steel was erected from January to June 1953. Most of the equipment came in from January to March of 1953 by tractor sleighs. Sane sleighs upset but no equipment was lest through the ice."

Jan 1, 1972

By J. L. Huiatt, R. B. Tippin

Introduction Great Salt Lake Minerals & Chemicals Corp. (GSL), a subsidiary of Gulf Resources & Chemical Corp., is located on the east side of the Great Salt Lake, approximately 20 miles from Ogden, UT. GSL processes brine from the north arm of the lake to produce potash, salt cake, salt, and magnesium chloride brine. Potash fertilizer, in the form of K2SO4 is GSL's primary product. The CSL Sulfate of Potash Complex: The GSL potash complex (depicted in [Fig. 1]) is divided into a solar evaporation pond system and the process plants. The solar evaporation system consists of a 17,000-acre (68.8 km2) ponding operation where the various salts deposit during the summer evaporation season and a harvesting operation which recovers the deposits during the remainder of the year. The harvest is either preconcentrated in the flotation plant or processed directly in the chemical plant where sulfate of potash (SOP) is produced. The chemistry of the lake brine results in a solar deposit containing primarily kainite (KC1 . MgSO4 . 3H20), schoenite (K2SO4 . MgS04 . 6H20), and halite (NaCI), with some carnallite (KC1 . MgC12 . 6H20) and epsomite (MgSOq . 7H20). Other minerals may be deposited, depending upon the weather during the evaporation season. Each year, the solar deposit in the evaporation ponds will greatly depend upon the summer weather conditions, which have in the past proven to be quite unpredictable. Weather may alter both the quantity of the solar deposit and the composition of this deposit. During a "short" evaporation season, the tonnage deposited decreases, but depending upon the short-term weather, the potash mineral content can vary greatly. During an "erratic" evaporation season, with intermittent summer rains or

Jan 1, 1980

Improve the Bureau-developed system for locating and communicating with miners trapped underground following a mine disaster. Approach Include in the wireless system a portable second-generation transceiver that is more reliable, compact, easier to deploy and operate, and ruggedized to withstand a mine environment. How It Works In an emergency, the new transceiver is snapped onto a caplamp battery for power. (Normally the 1 1/2-pound transceiver is either carried by the miner or is strategically located in the mine.) Its antenna wire, coiled inside the transceiver container, is removed and deployed in a roughly square configuration around a coal pillar or on the mine floor. The antenna consists of 300 feet of lightweight wire, which provides a large magnetic moment while remaining intrinsically safe, The photograph shows the

Jan 1, 1981

By Jaidev Sankar

The dream of constructing the new Second Avenue Subway in New York has finally become reality as the first construction contract package is currently underway. This first contract requires the construction a TBM Launch Box in the middle of Second Avenue from which a Tunnel Boring Machine (TBM) will bore two tunnels south to 63rd Street. This paper will discuss the design of this contract and how construction has progressed including preparation, site constraints, utility construction and com-munity impacts while highlighting the challenges of working in an urban environment. It will also discuss moving forward on this major project outlining the remaining contract packages including their scope and status.

By A. Sullivan

Success in mining demands the creative application of equipment, systems and processes to the unique needs of each mineral deposit, market and, though often overlooked, the culture of the individual business. In recent years, domestic coal mining companies have become more attentive to the contribution safety and loss control programs have on their financial results. Although most have adopted fairly conventional safety programs, some have created systems and processes tailored to their unique circumstances. While each coal mining business must consider the typical risks inherent to our industry, excellence in safety is most likely achieved when a safety program is customized to meet the needs of the individual business.

Jan 1, 2000

By S P. Young, K R. Monro

Reducing metal prices have forced mining operations to become more efficient and innovate in order to do more with the existing base. Much of this effort has focused on improving the efficiency of load and haul operations. The key to this is to understand and quantify where delays and missed opportunities exist. Metrics such as utilisation of availability, efficiency utilisation and asset availability are helpful indicators of performance; however, uptime-based time usage models and performance analysis tools using productive operating time models help to identify additional improvement opportunities.Phu Kham is a copper-gold mine operated in the Lao People’s Democratic Republic (Laos) since 2008 by PanAust Limited. The mine is a single large open pit of 450 m depth, running a fleet of 51 100 t class rigid dump trucks and 300 t and 200 t shovels controlled with the aid of a fleet management system in order to achieve ex-pit material movements of 56 Mt/a.Commencing in late 2014, a range of initiatives were implemented in order to improve the efficiency of the mine haulage fleet. These include hot seating of all operator changeouts via changeout ramps, split shifting of crews to minimise queuing delays, maximising blast volumes and minimising blast event days to reduce blast exclusion delay events including opportune refuelling of equipment during blasting events, inspections of trucks during refuelling events, implementation of a dedicated haul road services team to help reduce wet weather related delays and infrastructure development for haul path efficiency improvements.These and other initiatives have allowed the mine haulage fleet utilisation of availability (UofA) to be sustainably increased from 85 per cent to 97 per cent (see Figure 1), efficiency utilisation to increase from 85 per cent to 90 per cent, and productive operating time (POT) to increase from 73 per cent to 84 per cent. The corresponding increases in efficiency of the haulage fleet has allowed an equivalent reduction of eight haul trucks from the fleet to meet continuing life-of-mine production requirements.CITATION:Monro, K R and Young, S P, 2016. Truck haulage improvements at the Phu Kham mine – the journey to 97 per cent utilisation of availability, in Proceedings International Mine Management Conference, pp 339-346 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jun 22, 2016

The Mineral Industry of Canada in 1951 Continuance of the remarkable, successive increases in the value -of Canada's mineral production, chief basis of the nation's post-war expansion, was one of the highlights of Canadian development in 1951, when, according to the preliminary estimate is-sued by the Dominion Bureau of Statistics and reported in the BULLETIN for February, 1952, mineral output for the year reached a new high of $1,228 millions. This marked an increase of 17.5 percent over the previous year, which was the first in which the billion dollar mark had been reached, and an increase of 143.7 per cent over 1946, when production was valued at approximately $503 millions. While higher prices for the principal base metals accounted for a large part of the 1951 increase, quantity gains were made by copper, nickel, zinc, asbestos. lead, cement, and strikingly important. by petroleum, the value of which exceeded for the first time that of coal production. Gold was an exception to the general trend, being lower in both quantity and value than in 1950. For the first time nickel .came within $9 millions of surpassing-gold as the leading mineral product. Metals and Fuels The value of all metals, estimated at $733 millions, was 18.7 per cent more than in the previous year. Nickel advanced 11 per cent in volume and 34 per cent in value. Copper was up 2.4 per cent in quantity and 21 per cent in value, zinc rose 6.6 per cent in quantity and 35 per cent in value, and although the volume of lead produced was lower by 8 per cent, its value increased by 17 per cent. The tonnage of iron ore was 31 per cent higher than in 1950.

Jan 1, 1952

By R Bresca, A McCallum, J Scealy

The Endeavor Mine, located to the north of Cobar, New South Wales, was purchased by CBH Resources Pty Ltd in September 2003. The operation has been producing lead and zinc concentrate continuously for over twenty-three years. In 2004, a water management program was instigated to improve the raw water use efficiency of the mining operation. The water management program involved the establishment of a Water Management Committee, a review and quantification of water use across the site, investigation and implementation of efficient water use practices, and workforce awareness training. The program has seen the amount of water used per tonne of ore milled reduced from an average of 1.08 kL/t in the March 2004 quarter, to 0.68 kL/t in the March 2005 quarter. This paper outlines the Endeavor Mine water management program and may serve as a model for other operations that are planning to improve their water use efficiency.

Jan 1, 2006

By G. Bezuidenhout, N. C. Steenkamp

Buying centres are the first point of contact that determines whether ore will be considered as having sufficient provenance to be considered ethically sourced. The information that needs to be captured on the blockchain platform ranges from reports on in-loco inspections and audits of mining and purchase sites. This will guide passive and active monitoring of mining activities, trade-offs for identification systems, including biometrics for the registration of artisanal sellers, scrutiny on the selection of security measures and companies for buyer centres and stockpiling warehouses, evaluation of active and passive infrastructure security, and remote monitoring applications. Assistance with selection and training in geochemical testing instruments to determine the grade of ore material or concentrates and sorting, on-the-ground investigations, guidance on monitoring methods from stockpile and packing area along the transport route to the final destination are required.

Jan 1, 2023

By George H. Ashley

At the beginning of the war, about 13 years ago, a conference was called in Washington to lay plans for pooling the coals of the United States. A careful review of the various systems of classification then extant showed none well adapted to the needs of coal pooling. The system of Tidewater pools then adopted, using numbers to distinguish one pool from another, was based on typical coals of different localities. At that time, the writer undertook the preparation of such a practical classification. Aside from drawing on his experience of over 20 years in the coal fields of the United States, reaching from the Pacific Coast and the Rio Grande to Rhode Island, he first examined and carded all of the large collection of coals in glass jars made by and for M. R. Campbell. Next, he carded all of the descriptions of coals he could find in the U. S. Geological Survey bulletins and elsewhere, as well as practical tests of coals by the Navy and other organizations and then took account of every published coal analysis of the U. S. Geological Survey and U. S. Bureau of Mines. A study of this information, involving thousands of recomputations, led to the presentation of a paper on "A Use Classification of Coal" before this Institute at the Chicago meeting, September, 1919. That classification was unique (1) in a classification based on "standard coal" which included moisture, but used standardized ash, sulfur and nitrogen, (2) in using physical properties to distinguish the major classes or orders, (3) in recognizing British thermal unit values in the minor subdivision, (4) in the large number of classes recognized (thirty-six), (5) in proposing mineral names for all classes, (6) in proposing a letter code for the several classes for practical use, (7) in proposing a code for expressing what has commonly been called "grade" of the coal, the grade code to be combined with the rank code to designate and describe a coal. The writer still thinks that the principles then offered, though not necessarily the exact proposals, are valid and desirable in any practical classification. The actual classification of coal by rank was based on the ratio of fixed carbon to volatile matter for the high-rank coals—the well-known

Jan 1, 1930

By K. O. Bennington

The thermodynamic properties of two manganese silicates, pyroxmangite and fowlerite (zincian rhodonite), were experimentally determined by the Bureau of Mines. The standard enthalpies of formation were determined by hydrofluoric acid (HF) solution calorimetry, the low-temperature heat capacities by adiabatic calorimetry, and the high-temperature heat capacities by differential scanning calorimetry over the range 322 to 1,372 K. Data for the standard enthalpy of formation (AH298), the derived standard entropy (S298), and the standard Gibbs energy of formation (AGf 298) are as follows: Pyroxmangite [(Mn0.9554Cao,0446)(Si03)J- AHf298 319.668±0.46 kcal/mol AH298 (from oxides)... -7.215±0.11 kcal/mol 5295 23.785±0.048 cal/(mol'K) AGf298 -301.190±7.96 kcal/mol Fowlerite [(Mno.7o3Zno.0884Ca0.2086)(Sio3)]--AHf298 ?... -329.914±0.419 kcal/mol AH298 (from oxides)... -8.400±0.104 kcal/mol 5598 21.949±0.059 cal/(mol-K) AGf298 -310.716±7.99 kcal/mol These experimentally determined values were combined with data from the literature to calculate the Gibbs energies and equilibrium constants of formation over the temperature range of the measurements. Standard enthalpies, Gibbs energies, and equilibrium constants of formation are given as a function of temperature from both the elements and the constituent oxides.

Jan 1, 1987

By John Songstead

INTRODUCTION The Bunker Hill mine, located at Kellogg, ID, in the Coeur d'Alene mining district, has been a major pro¬ducer of lead-zinc-silver ore since 1887. Since its dis¬covery, it has produced over 25 401 171 t (28 million st) of ore from 20 to 25 veins and ore zones. The mine now covers an area of approximately 1.609 km (1 mile) in length and 0.8 km (1/2 mile) in width and extends from 1097 m (3600 ft) above sea level to 488 m (1600 ft) below sea level. Production is presently at a rate of 1542 t/d (1700 stpd). During 1965, a new stoping method, the Bunker Hill pillar system, was developed to enable the profitable mining of large, low-grade zinc ore bodies. The system was specifically designed to in¬corporate the Atlas Copco T26GH rubber-tired loaders. High productivities were achieved, and the method was adapted for mining the higher grade stopes where large timber square sets are utilized. This results in nearly 100% extraction. GEOLOGY The area encompassing the Bunker Hill mine lies immediately south of the Osburn fault, a major east-west striking, deep-seated strike slip fault which is the dominant structural feature in the Coeur d'Alene district. The Bunker Hill forms a series of veins and mineralized zones which occur as link structures, tension fractures, and breccia zones in folded, interbedded argillaceous quartzites of the Revett and St. Regis formations of the Precambrian belt series. The Rosco ore body is predominantly a zinc ore body above the Cate fault, which is presently known to occur from the 13 level. [elevation 488 m (1600 ft)] to the 1600 level [elevation 304 m (1000 ft) ]. These ore bodies are stringer zones with at least one dominant 1.8 to 4.5 m (6 to 15 ft) wide band of ore. They dip 0.09 to 0.26 rad (5° to 15°) steeper than the local 0.61 rad (35°) dip of the Cate fault, and also deviate 0.26 to 0.52 rad (15° to 30°) in strike from the Cate. Because of the flat dip and fractured weak hanging wall, normal cut-and-fill mining has not proven practical in this ore body. The mechanized timber methods described in the next section and the Bunker Hill pillar system are the mining methods successfully used. MINING METHODS While pillar stoping and mechanized cut-and-fill stop¬ing methods were obviously superior to the widely used timber stoping methods, as far as cost and productivity

Jan 1, 1982