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By A Fyans, B Ghahreman-Nejad

This paper discusses the method and results of an options study into selection of the optimal form for a 130 m high tailings/water retaining dam, located in a region of high seismicity in central Iran. The proposed dam is part of a planned expansion project for Sarcheshmeh Copper Mine (SCM), which after construction will provide 550 million tonnes of additional tailings storage over the next 27 years. Geotechnical investigation at the proposed site indicated that adequate suitable low permeability materials (i.e. clay) were not available in the immediate vicinity of the dam site and would need to be imported from a distance of well over 10 km (direct). Given the magnitude of the estimated quantity of material and construction cost it was considered prudent to compare the cost of alternative dam design options for the proposed storage. A total of five dam design options are considered including a conventional rockfill embankment dam with an upstream sloping earth core (ECRD), a geomembrane face rockfill embankment dam (GFRD), a sloping asphalt core rockfill embankment dam (ACRD), a roller compacted concrete dam (RCCD), and a filter rockfill embankment dam for tailings storage together with a separate smaller earthfill core rockfill dam (ECRD) located downstream for storage of decant water. The risks, costing and design of all options are discussed in detail and the outcome of the study is presented.CITATION:Ghahreman-Nejad, B and Fyans, A, 2018. Optimum embankment design for a tailings dam, in Proceedings Mine Waste and Tailings Stewardship Conference 2018, pp 427–438 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 23, 2018

By K. Fountain

In Novemeber of last year I visited a mica mine in Hunan Province, China. For nine months , we had been trying to sell Chinese mica to a customer in Belgium. Talking to the customer was like negotitating with a stone wall. The buyer refused to communicate any -thing other than his steadfast refusal to consider Chinese mica. He would not advise us what he wanted, what specifications or grades. He would give no indication if our price at market, or reasonable, or totally out to lunch. He would not even accept samples. It took no genius to guess that the mica buyer had at some time in the past bought a shipment of Chinese mica and had received a disaster. I invited the mica mine in Hunan in order to see for myself what could be done - without massive investment in equipment and instruments - to upgrade the quality of the mica for export. We had seen this type of customer reaction before. One disgruntled customer who had pur¬chased pigment grade barites direct from China ranted and raved about unreliable and inconsistent quality. It did no good to explain to him that China is a big country, with hundreds of mines scattered throughout many areas, each producing a different product. Our further explanation that Chinese export organizations also varied in degree and efficiency of customer service fell on deaf ears. Samples had no effect on his perception. Eventually we showed him photoraphs of two mountains of pigment grade barites. One lot was our sales of pigment barites to a major European minerals processor. Our Chinese barites stood proudly white and beautiful next to a hill of other material which now appeared at best mediocre and dull by comparison, and which the buyer had decided to reject after he received our Chinese barites. Finally the customer was sufficiently impressed to look seriously at our guaranteed specification and the outturn analysis performed by our other customer.

Jan 1, 1987

By M. M. Wong

Electrolysis of titanium carbide and a material consisting of titanium, carbon, nitrogen, and oxygen as soluble anodes in Used electrolytes was investigated by the Bureau of Mines for the preparation of crude titanium metal. Tests made in all-chloride electrolytes showed only a low recovery of titanium as well as a substantial volatilization of titanium trichloride (TiC13) when present in the electrolyte. Electrolysis at 750° to 850° C in a mixture containing 85 percent NaC1-15 percent potassium floutitanate (K2TiF6) yielded as high as 91 percent recovery of titanium from titanium carbide and 62 percent recovery of titanium from the titanium-carbon-nitrogen-oxygen material. Problems associated with the preparation of crude metal by fused-salt electrolysis are discussed. The crude metals produced from both anode materials were electrorefined in a sodium chloride-titanium chloride (NaCl-TiC12) electrolyte, and ductile titanium metal was obtained.

Jan 1, 1963

By O. M. Johnson

Barite and bentonite are the two principal components of drilling fluid systems used in the petroleum industry for oil and gas well drilling with rotary-type rigs. Approximately 87 percent of the world's annual production of barite is used in drilling applications. Of the worldwide production of bentonite, an estimated 30 percent is used. The drilling fluid, or drilling mud, to which barite and bentonite are added, performs five distinct functions: Lubrication-The liquid phase of the drilling mud lubricates the drill bit, preventing premature failure of the cutting edges and bearings. it also helps to soften the formation being drilled. Cuttings Removal-A mud system must be able to carry rock cuttings broken loose by the bit up to the surface where they can be removed by mechanical or chemical methods. The fluid's ability to perform this function depends upon maintaining a proper balance between the rate of circulation of the fluid and the colloidal properties required to entrain the cuttings in the mud to prevent their settling to the bottom of the hole. Hole Stability-The pressure exerted by the mud against the wall of the bore hole prevents the collapse of the uncased portions of the hole. Downhole Information-The fluid characteristics of the mud should be such that good cuttings samples, cores and electric logs may be obtained to help evaluate a well's productive zones and potential. Control Oil, Gas and Water Flows-It is essential that a drilling fluid offset any pressure from naturally occurring formation fluids in order to prevent a potentially wasteful and dangerous blowout.

Jan 1, 1980

By Kanchan Mondal, Anthony (Sam) Spearing, Gopi Bylapudi, Joseph C. Hirschi

"The U.S. coal mining industry uses about 100 million rock anchors per year. ASTM F432 specifies the requirements for rock bolts on mines but does not address corrosion. Corrosion has been found to be an issue in Australian coal mines (Hebblewhite et al., 2003 and Villaescusa et al., 2007), where the problem has been researched and stress corrosion has been found to be a significant cause of rock falls. Corrosion is also a major concern for underground civil construction in the U.S. but has not been considered an issue or adequately researched in coal mines. Underground conditions are conducive to corrosion mainly because of water quality and humid conditions. There is a perception, however, that when bolts are fully grouted, adequate corrosion protection is offered to the steel. Research has shown that this is not necessarily the case, due to the formation of micro-cracks as the resin sets and with subsequent rock movement shearing the resin column. This paper outlines a method to determine the corrosion potential of bolts used in long-term excavations and suggests ways to mitigate such effects based on research conducted in the lab and on three mines.INTRODUCTIONResearch on rock anchor corrosion in mines has not been extensive enough in the past, and the effects of several factors in the mine atmosphere and waters are not clearly understood. One probable reason for this lack of research may be attributed to the time required for gathering meaningful data, which makes the study of corrosion quite challenging.Almost all the rock bolts used in coal mines are made from steel, hence the corrosion of iron needs to be understood. The corrosion of iron is an electrochemical phenomenon with the overall equation"

Jan 1, 2010

By Christopher E. Dianora, Matthew Sorrell, Eileen Test

"Constructed between 1939 and 1945, the Rondout-West Branch Tunnel (RWBT) portion of the Delaware Aqueduct is 45 miles (72 km) long. The New York City Department of Environmental Protection (NYCDEP) first identified suspected areas of leakage from the RWBT in the 1990s. A $1 billion project is underway to mitigate the leakage through the construction of a bypass tunnel around the area of most severe deterioration, a 2.5-mile (4 km) section located beneath the Hudson River. Two shafts, 800 and 900 feet (244 and 274 m) below the ground surface, are being sunk to support the tunnel construction. The deep shafts will be subjected to high groundwater pressures and have been designed to accommodate multiple tunnel portals. This paper discusses the engineering decisions and analyses leading up to the final design. INTRODUCTION Project Background The New York City Department of Environmental Protection (NYCDEP) provides over 1 billion gallons (3.7 billion liters) of water to over 9 million residents of the City of New York and surrounding communities each day. This water is stored in upstate reservoirs and conveyed to the City via a system of aqueducts and tunnels. Water is delivered to the City almost completely by gravity and consists of three systems: the Croton, the Catskill, and the Delaware. The Delaware System is the largest of the three and supplies between 50% and 70% of the City’s daily water supply. In the 1990s, the NYCDEP discovered leaks in the central 45-mile-long (72 km) portion of the Delaware Aqueduct known as the Rondout-West Branch Tunnel (RWBT). The leaks release between 15 and 35 million gallons per day (MGD) (132.5 million liters per day) and are concentrated in two areas near Wawarsing and Roseton, New York. As part of the Rondout-West Branch Bypass Tunnel and Wawarsing Repairs Project (Bypass Tunnel Project), the leaks in the Wawarsing area will be repaired by grouting from inside the existing RWBT, and the leaks in the Roseton area will be mitigated by constructing a bypass tunnel parallel to and 1,750 feet (533 m) north of the leaking portion of the existing RWBT. Two deep shafts will be excavated on either side of the Hudson River: Shaft 5B in Newburgh and Shaft 6B in the Town of Wappinger."

Jan 1, 2016

By Gregory M. Rumbaugh, Todd R. Kostecki, Christopher Mark

The paper examines five massive pillar collapses in underground stone mines in the eastern United States from 2015 to 2021. These collapses caused significant airblasts, surface subsidence, and damage to mine infrastructure. The study identifies factors contributing to pillar collapses, such as slender pillars with width-to-height ratios of 0.8 or less, and proposes a framework for evaluating and mitigating collapse risks.

Nov 1, 2023

By R R. J McAllister, K Moffat, S Heyenga

This paper examines the statistical relationship between quality of life indicators and the gross value of mineral production from Australian regions. We used quality of life indicators, aggregated for 673 local government areas, of household income, housing affordability, access to communication services, educational attainment, life expectancy and unemployment. We find no evidence of systematic negative associations between quality of life and the gross value of minerals production. Instead, mining activity has a positive impact on incomes, housing affordability and communication access across regional and remote Australia. Whilst we do not establish causality between mining activity and quality of life, our analysis prompts a rethink of the æresource curseÆ as it applies within a single country. Nevertheless, our findings are tempered by extensive anecdotal and empirical evidence of the negative social impacts of mining on specific demographic sectors, localities, families of fly-in, fly-out mining operations and individuals. This contrast may be a scale issue, with the regional benefits of mineral wealth masking highly-localised inequalities and disadvantage. We suggest that there is a need to better understand these impacts and, more importantly, the types of policy mechanisms government and industry can adopt to mitigate or avoid them.

Jan 1, 2009

By Silas C. Nichols, Jennifer E. Nicks

"Limit states design (LSD), or load and resistance factor design (LRFD), is a method to account for the uncertainty in loads and resistance through a reliability calibration that requires all applicable limit states (strength, service and extreme) to be checked during foundation design. LRFD has been in use internationally for several decades; however it wasn’t until recently that engineers working on highway projects were required to use the method for the design of highway bridges and other appurtenances. Although the bridge community has been using a form of limit states design for structural components since the early 1970s, the design of foundations for bridges, walls and other geotechnical features had long been performed using allowable stress design (ASD), whereby all uncertainty in loads and material resistance is combined in a single factor of safety. While the move to LRFD was viewed as a more rational approach for the design of bridges and structures, the transition to its use for geotechnical engineering has been met by many with confusion and difficulties in application. This article will revisit the history of transition from ASD to LRFD for geotechnical engineering, summarize the status of the transition and issues with its use, and briefly discuss some emerging research opportunities for improving and advancing the design platform.Background and HistoryUntil about a decade ago, foundations and substructures for bridges were traditionally designed using ASD, while the superstructure was designed using load factor design (LFD), a precursor to LRFD that incorporated factored design loads but no risk assessment. This combination led to uncertain and incompatible safety margins in design for the various bridge components. In 1986, the American Association of State Highway Transportation Officials (AASHTO) Subcommittee on Bridges and Structures (SCOBS) concluded that the standard bridge specifications in use at the time for U.S. transportation projects had significant inconsistencies and gaps, wasn’t up to date with emerging technology and that the development of new specifications was warranted. The proposed solution by the subcommittee was to transition to a more coherent method of design where applicable failure and serviceability conditions could be evaluated considering the uncertainties associated with loads and resistance."

Jan 1, 2015

By James V. Luxner

The airflow and methane distribution patterns developed in an equipment-free entry 6.5 feet high and 12 feet wide, ventilated by line brattice operated blowing to or exhausting from the face, were determined by the Bureau of Mines for varying conditions. Results of these studies show that blowing line brattice systems of face ventilation are more effective than comparable exhaust systems. The face ventilating efficiency of any line brattice system is dependent upon the method of ventilation, the face distance, the tight rib distance, the volume of air delivered to the end of the line brattice, and the volume of methane released at the face. Experimental results indicate that maximum face ventilation efficiency is achieved when the inby terminus of the line brattice is installed 5 feet from the face and at a distance less than one-third the entry width from the nearest rib. Analyses of data are based upon the ability of a face ventilation system to dilute methane in the immediate face area.

Jan 1, 1969

By Baruch Grinbaum, Edwin Bwalya, Robert Minango, Deogratias Bukunkwe

Solvent extraction of Cu from acidic leach, using an oxime extractant, e.g. LIX 984 or Acorga is a well known and established process. As the process advances, Cu is extracted to the solvent and is replaced by H+ in the aqueous phase. At high acidity the distribution coefficient of Cu between the solvent and the aqueous phase decreases, and eventually relatively high concentration of Cu remains in the raffinate, regardless of the number of equilibrium stages. In many processes this phenomenon is immaterial, as the raffinate returns to the leaching and no copper is lost. But when the copper raffinate is used for processing of additional metals, minimal concentration of Cu in the raffinate is essential for an effective process. The recommended process includes three steps: 1. SX-1: Standard extraction of Cu with oxime. The resulting raffinate contains about 6-8 g/L Cu and 50 g/L H2SO4. 2. SX-N: Extraction of most of the acid into an amine-based solvent, and stripping it with a base, e.g. aqueous ammonia, caustic, Na2CO3 etc. The extraction removes no metals, but decreases the concentration of acid from 50 to 10 g/L. 3. SX-2: Extraction of the neutralized solution once more with oxime, to remove the residual copper. Such a process enables one to get down to 100-150 ppm Cu in the final raffinate. This process was developed and tested in the R&D facility in Chambishi Metals plc, by a joint team of Chambishi & Bateman, using Cognis LIX 984N and real solutions from the plant. In the laboratory scale, the equilibrium isotherms for SX-1, SX-2 and the neutralization were developed. Computer simulation of the results showed that the optimal configuration is two stages for both extraction and stripping for SX-1. For SX-2 2-3 stages of extraction and two stages of stripping are needed. The stripping stages may be common to both SX-1 and SX-2. The neutralization of SX1 raffinate was tested both by conventional precipitation with lime and by extraction with Alamine based solvent extraction and stripping with a solution of 10% NaOH. For the SX-N only one stage was needed for both the extraction and stripping. The bench scale pilot plant was run for both SX-1 & SX-2. The results were very similar to the simulated ones, and the configuration of 2E2S for SX-1 and 3E=2S for SX-2 delivered the required raffinate with less than 150 ppm Cu.

Jan 1, 2005

By Matthew N. Rubin

This report summarizes the work performed under USBM Contract No. H0155155, "Demonstrating the Noise Control of a Coal Preparation Plant." A Consolidation Coal Company coal preparation plant (the Georgetown Plant) was selected for demonstration of various noise control techniques and materials in a commercially operating plant. At the start of the project, the 1650 tons per hour (tph) plant was surveyed to document the existing sound levels and worker noise exposures. On the basis of the survey, individual noise sources were characterized and assigned priorities for treatment. Four categories of noise control treatments were installed in this plant: ? Resilient impact pads ? Curtains ? Chute liners ? Resilient screen decks. Currently (1977) available resilient impact pads and curtains were found to be suit-able for general use in coal preparation plants; in the Georgetown Plant, they provided significant noise reduction [6 to 10 dB(A)]. The overall performance of the chute liners which also provided useful noise reduction, was strongly dependent on the specific installation. Although the resilient screen decks have the potential for 7 to 9 dB(A) of coal/screen impact noise reduction, blinding and delamination (in composite decks) en-countered at the Georgetown Plant represented major design problems for these products. This program achieved noise reductions of up to 10 dB in various areas of the Georgetown Plant (within the restrictions imposed by the plant's operation) resulting in significant reduction in worker noise exposures.

Jan 1, 1977

By C Donnelly

Redpath Australia is using a unique mining method for the Australian underground coal industry at Kestrel Mine Extension in the Bowen Basin, Central Queensland.The project involves completing two separate drifts from the surface for access to the German Creek Coal Seam located approximately 230 m below the existing surface. The scope of work is primarily the construction of two drifts; the first is 1577 m at 1 in 6 gradient and the second 1874 m at 1 in 8 gradient. Each drift has been developed by a Mitsui S200 roadheader with modifications to allow for a continuous tunnelling/mining cycle operation. A shotcrete boom has been fitted to the left hand side of the machine with the operators console located at ground level. Another operators console is elevated on the right hand side of the machine to allow operation of the rotary drill boom which is mounted on a slide rail to allow flexibility of positioning depending on the stage of the cycle. In support of the roadheader activities, an integrated conveyor system mounted on a sliding floor arrangement is able to be advanced as the mining face progresses. This sliding floor consists of four 1.5 m × 6 m sections that utilise hydraulic rams to ‘walk’ inbye, allowing room for installation of precast concrete floor panels for final floor completion. This method borrows from the tunnelling and construction industries to eliminate place-change delays in the mining cycle by enabling complete ground support installation without retreat of the primary development machine from the face. The aim is to reduce the project duration with the completion of final floor and services to within 30 m of the operating face as it advances.

Mar 21, 2011

By R. Drozd

"The Bui/moose Mine is in the northeast coal area of British Columbia, and less than $300 million was spent to bring this 2.3 million-tonnes-per-year metallurgical and thermal coal mine into production by November 1, 1983, one month ahead of schedule. The project is an unincorporated joint venture undertaking, with Teck-Bullmoose Coal Inc., a wholly-owned subsidiary of Teck Corporation, holding 51%, Lornex Mining Corporation holding 39%, and Nissho-lwai (Canada) Ltd. holding 10%. Bullmoose Operating Corporation will be the manager and operator of the project. This paper will broadly cover the history and exploration work on the property, the geology, open-pit mining plans and equipment, the preparation plant flowsheet from breaker station to clean coal silo, the 35 km highway truck haul, the railroad loadout system, manpower requirements, approvals and permits, and general financial arrangements. Introduction The Bullmoose Mine Project is located about 87 km by road south of Chetwynd, B.C. and consists of an open-pit coal mine, a plant facility in the Bullmoose Creek valley below the mine, and a separate rail loadout facility on the British Columbia Railway Branchline near the new town of Tumbler Ridge some 35 km by road to the east. The plant will be capable of producing a total of 2.3 million tonnes per year (mty) of washed metallurgical and thermal coal. Long-term metallurgical coal contracts for high-quality, medium-volatile, bituminous coal have been signed with nine Japanese buyers for 1.7 mty, and the balance of0.6 mty will be for thermal coal production. Plant construction began in the spring of 1982, and in May 1983, preproduction waste and overburden removal in the mine was started. The plant began tuning up in October and was fully operational by December 1983. Metallurgical coal began moving to the coast in unit trains during November, and first vessel loading at the Ridley Island port began in late December 1983."

Jan 1, 1985

By Tim M. Benson, Adam L. Lamanna, Carolina Martinez-Gutierrez, Mark J. Stephani

The Lower Concourse C Extension (LCE) 108-in Storm Sewer consists of a 2,797-foot long, 27 to 55-foot deep, 108-in I.D. storm sewer tunnel. Situated beneath active taxiways and a terminal apron at Hare International Airport (D), the tunnel diverts ows from existing storm sewers located within the footprint of a future satellite concourse. The project was delivered via Construction Manager at Risk (CMAR) as part of Chicago Department of Aviation’s (CDA) ongoing O’Hare 21 Capital Program. This paper presents project details, multiple stakeholder coordination, subsurface risks and challenges, TBM performance, and impacts to existing infrastructure.

Jun 23, 2024

By S. Banisi, E. Arghavani, Gh. Parsapour, S. Mousavi

Maximization of water recovery in thickeners is an essential step in the mineral processing industries. The Gol-E-Gohar iorn ore processing plant uses a center pier thickener with a diameter of 38 m and depth of 4 m. The feed rate is 50 t/h solids with a concentration of 5% (w/w), and it has been designed to provide an underflow containing 50% solids. The performance of the thickener was not as expected judged by the overflow stream containing 0.6% solids and water loss of 52 m3/h in the underflow. A criterion called the index of flocculation incorporating the perimeter-based fractal dimension of flocs, floc size, and settling velocities was proposed and used to determine the optimum flocculant dosage which was found to be 20 g/t. The fractal dimension of flocs in a pulp containing 5% solids was found to be between 1.16 (no flocculant) and 1.56 (40 g/t flocculant) which indicated highly branched and loosely bound structures for the flocs as flocculant dosage increased. At these conditions, the size of the flocs which was recorded by a digital microscope increased from 93 µm to 403 µm. A new feedwell with a special design with the aim of preventing flocs breakage and maximizing energy dissipation of pulp was constructed and installed. These modifications led to an increase in the thickener drivehead torque from 5% to 18% of the rated torque, indicating a higher bed depth. Finally, installation of a PID controller reduced the drivehead torque relative oscillation from 52.0% to 12.7%. These changes increased the solids concentration of the underflow from 49.0% to 56.9%, decreased the overflow solids concentration from 0.6 % to values smaller than 0.01% and resulted in a reduction of 27% in water loss in the thickener underflow.

Jan 1, 2014

By Milton H. Fies

THE writer wishes to acknowledge at the outset his great sense of obligation to those who contributed so broadly and expertly to the preparation of this paper: Dr. Albert DeSmaele, Chairman of the Board of Directors of Socogaz, Belgium; Dr. Julien Venter, Director of the National Institute of Coal Mining and Coal Utilization Industry, Belgium; Dr. Roger Loison, Chief Engineer, French National Coal Research Board, France; and Dr. Erich Sarapuu of the Sinclair Coal Co., Kansas City, Missouri. Liberal use has been made in this paper of a recent publication on Russian, Italian, Belgian, and French experiments prepared jointly by the Centre d'Etudes et Recherches des Charbonnages de France (Cerchar) and the Institut National de l'Industrie Charbonnière of Belgium (Inichar).1 The writer is particularly indebted to Dr. E. T. Wilkins, of the Fuel Re¬search Station, London, and to Mr. James L. Elder, of the Alabama Under¬ground Gasification Experiment at Gorgas, Alabama, for their unfailing patience and effective support. Not only did these experienced and capable men edit and revise this discussion, but much of its content is a direct con¬tribution quoted in full from their memoranda. Without the assistance of all these colleagues, this presentation would not have been possible. It has been a privilege to serve and work with them and to acknowledge repeatedly their loyal cooperation. The recovery of the energy inherent in coal beds throughout the world has been done through the years by methods that today may be justly regarded as primitive. While the techniques of coal mining through the application of machines and the enhancement of information and skills have advanced to an extraordinary extent, particularly during the past 20 years, the underlying means is essentially the same as when man first began to mine coal. Despite this progress in the manner of coal mining, it continues today as a hazardous, toilsome, grimy task. Coal mining per se has advanced at a greater and more rapid stride in the past 25 years than the mining of any: other mineral, but the basic method is fundamentally the same as it was the first day a miner took a pick into his hands many centuries ago. This situation presents a challenge to mining engineers everywhere, from

Jan 1, 1953

By Paul Staples, Paul Messenger, Greg Lane

"Every project is different. While “conventional” SAG and ball mill circuits will suit many projects, for others there are alternatives, which may produce improved outcomes. This paper examines particular examples of recent front end designs which either went against conventional thinking or which employed “out of the box” thinking to realize the desired results.One such example is the design strategy employed for a small capacity gold plant in Victoria, Australia. The selection of HPGR technology for small capacity projects is not generally favoured, given the relatively high cost of small HPGR machines. However, the unusual size distribution of the gold and its liberation characteristics were expected to lead to improved gold recovery offsetting the higher capital cost.Another example is the first large scale application of milling in acid for copper recovery, at a project in south east Asia. The stainless steel, single stage SAG mill using ceramic grinding media was more expensive than a mill constructed of conventional materials. However, the steps of scrubbing, slurrying, grinding and leaching could all be achieved in a single unit operation and the need for thickening and repulping with acid after milling was eliminated.Rod mills were selected for both primary and secondary milling duties for another south east Asian project despite their somewhat outdated reputation. A review of previous industry experience confirmed that rod mills could outperform conventional rod/ball or SAG/ball circuits in cases where the size distribution of the ground product is critical to downstream recovery.These examples are discussed with particular focus on the key drivers in each case."

Jan 1, 2008

By R D. Lama, A Misra

One of the major factors restricting an increase in the productivity of a longwall mine is lower productivity obtained in a development panel. There has been a significant improvement in gate road development, particularly in mines requiring high density roof and rib bolting within 2 m of the face. Advance rates have increased from an average of 3 m to 5 m per shift in 1985 up to an average of 5 m to 10 m per shift in 1991. In mines at shallow depths with reasonably good conditions the increase has not been that significant. Advance rates have increased from an average of 7 m to 12 ni per shift in 1985 to an average of 12 m to 15 m per shift in 1991. Compared to the United States' (US) development rates, these are much lower. Extensive research work has been carried out to minimise cutting, wheeling, support setting time, and to optimise pillar length to achieve maximum advance rates. Pillar optimisation is very important in order to expedite adverse rates. This paper investigates various development rate technologies, both present and future, and optimises pillars to improve these dev elupment rates. Pillar optimisation is based upon time and motion studies conducted at mines in the Illawara region of New' South Wales, Australia, and the basic tools of arithmetic.

Jan 9, 1996

E. E. Russell Tratman, New York City (communication to the Secretary): Mr. Glenn's paper brings up again the question of the part played by coal-dust in colliery-explosions. In the discussion of Sir Frederick Abel's paper on " Accidents in Mines," read before the Institution of Civil Engineers,? much was said about the dangerous properties of coal-dust, and several instances were given of the successful use of a spray of water, water and compressed air, or exhaust-steam from an engine, to keep the mine SUEciently damp to prevent the formation of dust clouds. Methods of preventing shots from blowing out; of using water cartridges to prevent flame ; and of using lime cartridges, wedges and other means of breaking down the coal were also discussed. Mr. Emerson Bainbridge referred to the explosive termed " roburite " (ammonium nitrate and

Jan 1, 1895