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This report summarizes a study conducted by The Service Bureau Corporation for the Bureau of Mines. Evaluations reflected herein are the result of an independent investigation by The Service Bureau Corporation of methane considerations during coal mining. The study is oriented toward a systematic analysis of methane control problems during coal mining. Throughout this study an attempt has been made to ensure considerations consistent with the recently enacted Federal Coal Mine Health and Safety Act of 1969. References to reports by The Bureau of Mines and others are made for the purposes of establishing identification between the evaluations contained in this report and on-going research efforts.

Jan 1, 1970

Resources and reserves are critical inputs into the determination of an operating mineÆs financial results and balance sheets, but how well understood is this fact. Do the investors, directors, accountants and geologists/engineers responsible for preparing or relying on relevant aspects of the end results properly understand the inputs and outputs and the associated risks that may be established. The first step is to understand how financial reports are impacted. The most obvious, and best known, impact is in the calculation of depreciation and amortisation charges against long term assets. In addition, however, there is a requirement to regularly determine whether asset book values are impaired û this is nearly always performed globally by the preparation of a discounted cashflow model for a mine which is usually based on a mine life as determined from the resource and reserve statement û the particular driver here is normally the short and long term production plans. A key aspect of the short term production plan is that it would normally not assume that the reserve and resource statement implies consistent grade, stripping ratios, etc. However, outside of the early years of a plan, more generic assumptions of these variables are often made in conjunction with the reserve and resource statement. Depending on the timeframe of the detailed production plans, this can have a large impact on a projectÆs net present value. Other financial balances that may also be impacted include restoration and rehabilitation provisions, retrenchment provisions and deferred or equalisation stripping accounts. The age old question of whether all reserves should be used or whether some or all of the resources should be included for financial reporting purposes has never been adequately resolved. It is our view that there should be no prescriptive answer to this question but judgement of all professionals involved in the process needs to be included. Furthermore, the company needs to be able to support their conclusions with well thought out and appropriate answers. This may not be in line with the approach that certain accounting standard setters may be proposing or taking, and this paper will outline the current status of global requirement and the likely future directions. Whether reserves or resources are utilised, does not change the key requirement that those preparing or using reserve and resource statements need to understand key aspects. For example: geologists/engineers do they know how reserves or resources will be used? have they imparted information on variability of key drivers such as grade or depth or dilution? do they know that major financial/business altering decisions may be made based on the analysis? accountants do they understand where the numbers have been derived from and the assumptions that may have been made in their determination? do they understand the degree of accuracy (or inaccuracy) that may exist particularly with resources)? do they understand the history of resource reconciliations and what it may mean for the future? directors do they understand the processes undertaken by the company to prepare the financial report and the interaction between the geologists/engineers and accountants? do they understand the potential sensitivity of results to the input variables? Financial reporting is extremely reliant on the use of the reserves and resources, however, the geologists/engineers, accountants and directors do not always understand the interactions between themselves which may lead to misleading public reporting. More education and communication between various parties is needed.

Jan 1, 2002

By G. Boehler

The purpose of this program is to demonstrate the feasibility and practicality of noise control for secondary and tertiary crushers in surface non-coal mining plants and mills. It is a part of the broad Bureau of Mines program to reduce noise in mines to levels mandated in the Coal Mine Health and Safety Act. Utilizing the results of previous contractual research efforts in which noise levels of secondary crushers had been measured, two noise enclosures were designed and evaluated for the REXNORD Symons 7-ft diameter secondary and tertiary crushers installed at the Jackson County Iron Company, in Black River Falls, Wisconsin. The first enclosure had a circular shape following closely that of the crusher and was built up entirely of quick-release panels which can be handled by the crusher operator alone. The second enclosure was octogonal in shape and can be rapidly removed, partially or totally using an overhead crane. Both enclosures fully satisfy the noise reduction objectives and permit daily and long-term maintenance of the crushers. Cost analyses show the cost of the enclosure to be less than 10 per cent of the cost of the crusher, the cost of the octagonal enclosure being estimated at only 60 percent of that of the circular one, e. g. $15,000 against $25,000. It is concluded that the octagonal enclosure best meets the needs of the Jackson County Mining Company and it is recommended that an experimental enclosure be built at the earliest possible time, as proof-of-concept.

Jan 1, 1977

By B. Papillon

There has been a veritable increase in the adoption of electronic detonators in the last few years. Electronic detonators offer several distinctive features over electric and non-el (shock tube) detonators including better timing accuracy resulting in excellent fragmentation and vibration control, certainly key mining parameters. The field programmability also helps in reducing the on-site magazine storage of detonators with diverse fixed delays. A typical electronic detonator may consist of a logic core, a communication interface, storage capacitor(s) and the ignition element. Being electronic components, the electronic detonators may be susceptible to RF or electromagnetic interference (EMI). RF sources are virtually everywhere: cellular phones/towers, radio transmission, microwaves, walkie-talkies, high voltage (HV) transmission cables and wireless networks with their characteristic frequency emissions. Safety issues prevail; e.g., whether the RF interference can cause premature or inadvertent firing whether directly or indirectly. From a reliability perspective, the RF interference may also impact on bus communications, e.g., during the logging, verifying, charging and firing phases. The deleterious communications disturbance may result in a misfire (non-firing of one or more of the detonators) which is terribly inconvenient for the mine operators and even pose a safety hazard.

Jan 1, 2009

By Joe Holmes, Peter Langlois

"Falconbridge’s Raglan Concentrator, located at the northern limit of Quebec’s Nunavik Region, has been in operation since December 1997. Originally designed to process 800,000 tonnes per year of high grade nickel ore, plant capacity is currently approaching 1,000,000 tonnes per year. Optimization of the ABC grinding circuit has proven to be the most challenging component of this capacity increase. Concurrent flotation circuit modifications have allowed concentrate grades to increase from 16 to 18 percent nickel, while maintaining recoveries. Concentrate drying problems encountered during the first 15 months of operation have been addressed through equipment upgrades. The challenges associated with surface deposition of filtered tailings in an arctic environment are now better understood, and detailed design of the progressive reclamation cover is underway.GRINDINGCircuit OverviewOre from open pit and underground mining operations is introduced to the underground ore handling system via surface and underground truck dumps. Two underground bins with a combined capacity of 4000 tonnes provide roughly a day of live capacity for ore storage. Ore from the two bins is conveyed to a 150 tonne surge bin immediately ahead of the grinding circuit in the concentrator.The Raglan grinding circuit has a conventional ABC configuration (Figure 1) with a design capacity of 100 mtph (800,000 tonnes per year). The 24’ AG mill is powered by a 3000 hp motor with variable speed drive. AG discharge is sized on a 8’x16’ single deck vibrating screen, with screen oversize recirculating to a 5 %’ short head cone crusher (400 hp) before returning to the AG mill. Screen undersize is combined with ball mill discharge to feed up to 5 15” Krebs hydrocyclones. The ball mill is a 14’x 21’ overflow mill driven by a 3000 hp motor at 73% critical speed. Target Pgo for the grinding circuit product is 65 microns"

Jan 1, 2001

TABLE TUNNEL NAME OWNER LOCATION STATE TUNNEL USE LENGTH (FEET) WIDTH (FEET) BID YEAR STATUS Gateway Tunnel project Amtrak Newark NJ Subway 14,600 24.5 2015 Under study

Sep 1, 2013

By Walter G. Kutschke

Four significant soil nail wall projects in the eastern United States were recently completed with a combined area of 175,000 square feet of finished shotcrete surface. These projects required the use of innovative design and construction methods in order to address various challenges, including slide-prone back slope materials, perched water, highly erodible rock materials, curved wall alignments, very tight construction tolerances and unexpected subsurface conditions. Special focus is given to issues such as bench excavation, soil nail installation methods, shotcrete mix design, anticipated shotcrete quantities, shotcrete nozzlemen qualifications and weather conditions, in order to provide a lessons-learned database for future soil nail wall design and inspector considerations. These projects also underscore the importance of design engineer and soil nail wall contractor qualifications as well as effective communication with the owner.

Jan 1, 2007

By Shigeru Nishitake

The Earth Pressure Balanced (EPB) shields designed by Mitsubishi Heavy Industries, Ltd., are recently more enthusiastically accepted by tunnel boring specialists. This company having built 1,000 tunnel shields utilizes its vast experience to resolve the problem to excavate through big boulders scattered in rather soft alluvial soil. Several types of EPB shields which can cope with large size boulders are described. Such shields are equipped with both drag bits and rolling disc cutters on the cutterheads. Previously removal of boulders has presented safety problems for workers while in the muck chamber, or venturing in front of the shield. This report describes the problem and how the equipment was modified to deal with the problem.

Jan 1, 1987

By Stéphane Bourg

The SCRREEN project gathers more than 50 European initiatives, associations, organisations or independent experts working on Critical Raw Materials (CRM) into a long-lasting network including stakeholders, public authorities, and civil society representatives. SCRREEN contributes to improve the CRM strategy in Europe by (i) mapping primary and secondary resources and substitutes of CRMs, (ii) estimating the expected demand of various CRMs in the future and identifying major trends, (iii) providing policy and technology recommendations for actions improving the production and the potential substitution of CRM, (iv) addressing specific waste electrical and electronic equipment (WEEE) and other End-of-Life (EoL) product issues related to their mapping and treatment standardization, and (vi) identifying the knowledge gathered over previous years and making the data accessible to those beyond the project. As a first step, SCRREEN worked at providing stakeholders with a comprehensive analysis of the current use of the CRM, the mapping of potential primary and secondary resources, and identification of the key technologies used in their production. Substitution profiles, analysis of European business and policy issues pertinent to CRMs, and an identification of the standards, policies, and regulatory frameworks concerning these CRM was also completed. In a second step, to be conducted between mid-2018 and mid-2019, SCRREEN will provide guidance for improving the CRM market in Europe by identifying opportunities from research, development, and innovation (R&D&I) that could reduce the supply and/or economic risks around the most relevant CRM for Europe.

By Derrick D. Dasenbrock

Three piers of a new river bridge are supported on 1.067 m diameter open-end steel pipe piles. At time of drive, and after pile re-strikes, high-strain dynamic capacity values at plan tip elevations, were as low as 23% of design requirements. Pile lengths were extended 20% to 40% deeper than planned at two piers, driving to rock, impacting both the project cost and schedule. The significant discrepancy between predicted and observed capacity was investigated through a series of coordinated dynamic tests; a full-scale test pile was installed and monitored over 221 days by both dynamic and quasi-static methods. Unusually slow strength gains were observed by dynamic monitoring on all piles installed at the site. CPT soundings were advanced at the test pile and through the pile interior to investigate the soil plug. High-strain dynamic methods were found to have significant limitations in properly assessing the long-term capacity of large-diameter open-end pipe piles during driving and short-term pile re-strikes. As the benefits of large diameter high-capacity open-end pipe piles are numerous, the geotechnical community needs to address appropriate construction monitoring and control of these foundation types to take full advantage of these systems. An improved understanding of the physical effects of pile installation and the impact on soil strength assessment while driving and long term soil strength gain should be developed and incorporated into both design and monitoring procedures to avoid either unsafe or needlessly costly designs.

Jan 1, 2006

By R. Tessier, C. R. Larsen

"INTRODUCTIONOn November 10, 1982 Nor and a Inc signed an agreement with Golden Sceptre Resource Ltd. and Goliath Gold Mines Ltd. to construct a mine and milling facility.The Golden Giant mine is located in the Marathon area of north - western Ontario, immediately north of Highway 17 and about- 3.5 km west of Highway 614.With the exception of selected unit processes, the mill was designed for a nominal throughput of 2500 tonnes per day with early operating data to be used to identify sections requiring possible expansion to achieve an ultimate throughput of 3000 tonnes per day.Commissioning of the highly instrumented milling complex which includes crushing, grinding, cyanidation, and carbon in pulp processes commenced in February 1985. After water testing various sub-circuits from grinding through tailings, waste ore was introduced to ensure proper mechanical and electrical operation of process equipment.The first ore was milled as a mixture with waste on March 4, 1985. The first gold bullion was poured April 6, 1985.Initially, ore was only available from mine development at a rate of 700 t/d, which was milled at a minimum rate of 1200 t/d. While regular weekly shut downs of two to three days was useful for mechanical maintenance, changes, and modifications, process tuning was more difficult due to the non-steady operation. As increased quantities of ore became available, regular weekly shutdowns tapered off and process operation improved accordingly. For testing purposes and as ore was available, mill throughput was increased regularly to 2300 t/d and for very short periods of time to nearly 3000 t/d.No major design or operating problem has been encountered to date. The following is a brief description of the milling process and various operating highlights. Process flowsheets are shown in Figures 1 through 6."

Jan 1, 1986

By Ross W. Smith

Microflotation and surface chemical studies were conducted on a chromite from the Stillwater Complex in Montana and on some of the minerals usually found to be associated with such chromites. Considering the chromite it was found that aging the ground mineral in air or heating it markedly affected its flotation behavior. Much of the observed effects could be attributed to oxidation reactions involving surface chromium and iron species. Strongly acid pretreating the mineral appeared to reverse such effects. Thus, it can be concluded that the manner in which the mineral (or the ore) is pretreated is of utmost importance in its flotation. Work carried out on associated gangue minerals determined how they are likely to behave in a flotation environment and how certain foreign ions derived from a chromite ore itself can affect their surface characteristics and, thus, presumably also their behavior in a flotation circuit.

Jan 1, 1981

By Paul J. Axtell

Two new railroad bridges in Kansas City, Missouri are founded on drilled shafts embedded into the soil and underlying shale. The uncertainty arising from the non-homogeneous, anisotropic nature of most shales, and the variability of shales across the U.S., makes it a difficult material to assign appropriate design properties. As was the case on this project, the uncertainty of foundation performance in shale can lead to overly conservative designs, particularly with drilled shaft foundations. An extensive Statnamic axial testing program was performed to verify that the load-settlement behavior of the drilled shafts would be acceptable for the intended railroad traffic. Axial Statnamic testing was applied in three increasing cycles of 927 kips, 1007 kips, and 3177 kips, respectively. The maximum movements recorded at the shaft head during each of the loadings were 0.052 inches, 0.062 inches, and 0.257 inches, respectively. The loads applied in each of the respective Statnamic tests were not sufficient to reach the capacity of the shaft; so direct comparison of measured and calculated capacities is not possible. However, evaluation of the estimated design capacity can be accomplished based on results from the Statnamic testing program by utilizing normalized load transfer relations presented by O?Neill and Reese (1999). Such comparisons necessarily require extrapolation of the data measured in the Statnamic tests following typical load-displacement response. Results indicate the Kulhawy and Phoon (1993) method for estimating side resistance in Pleasanton shale rock sockets, assuming a smooth socket, appears to agree well with observed behavior during Statnamic testing. Accordingly, the O?Neill and Reese (1999) method for estimating the toe resistance in intermediate geomaterials, assuming cohesive rock with Rock Quality Designation (RQD) between 70 and 100, appears to agree well with observed behavior during Statnamic testing. Finally, based on Statnamic testing at varying load magnitudes and normalized curves developed by O?Neill and Reese (1999), the contribution of side resistance to the capacity of a deep foundation can be determined assuming no load is transferred to the base of the foundation during lower magnitude testing. Furthermore, higher magnitude testing can provide a means to evaluate the accuracy of design end bearing calculations.

Jan 1, 2006

By Fred D. Wright, Stefan H. Boshkov

12.1-BASIC AND PARAMETRIC CRITERIA IN THE SELECTION, DESIGN AND DEVELOPMENT OF UNDERGROUND MINING SYSTEMS STEFAN H. BOSHKOV AND FRED D. WRIGHT The criteria governing the selection of an underground mining system relate to several factors the importance of which varies with geographic location, the state of development of the industrial technology, the economic climate in the Nation and, perhaps, in the world-all these as functions of time. Thus, the best choice of a mining method depends upon sets of definable and undefinable values for numerous factors that may be considered fixed at the precise point in time. Subsequent changes in economic and/or technological developments may necessitate modification or complete revision of the existing mining system. A classification of criteria for selecting a mining system must be based on the broadest definitions. Specific cases may present simple solutions or pose very intricate questions, the answers to which are frequently based on judgment rather than explicit logically reasoned and derived results. Specific criteria may be of negligible consequence in some instances and of overriding importance in other cases. It would be foolhardy to attempt to point out, in the limited space available, all possible variations and interplay of factors, as well as to define the setting in which this interplay can take place. The reader is, therefore, referred to the case histories described in the subsequent parts of this section where actual practice will point out the variability of the problems encountered in the selection of a mining method. A popular thesis suggests that the proper approach to selecting the method of mining should be rooted in the inverse solution, whereby the limitations of the various methods in use serve as criteria to eliminate most of these methods from consideration in the specific case. This narrows the ultimate selection to one or perhaps two most appropriate methods. At this stage, a critical comparison analysis will lead to the ultimate choice. Although such a negative approach un¬questionably is valid in discarding obviously inapplicable mining systems from consideration, it falls short in defining the best system, which in practice quite often turns out to be a variation of a textbook standard, or a combination of two or more such standards. Because the deployment of capital and manpower over a long-term commitment closely follows the selection of the mining system, the decision-maker can well afford to spend a great deal of time and effort in evolving the optimum solution for the peculiarities of his specific case. Within the basic engineering tenets of economy and safety to which the design of a mining system must adhere lie numerous factors of constraint which may be grouped, somewhat arbitrarily but for ease of discussion, into several categories. The criteria that enter into the choice of a mining method derive from the following groups of factors: 1. The shape, size and spatial position of the ore deposit.

Jan 1, 1973

Splashing and wave formation has been studied experimentally by using one-seventh water model of CLU (Creusot-Loire Uddeholm) reactor with two different tuyere configurations by using paraffin and water as simulated slag and simulated metal respectively. Experimental results have been discussed in terms of total splash volume, relative slag volume in splash (dispersed-slag phase holdup in splash), wave frequency and wave amplitude with respect to gas injection rate and tuyere configuration at bath heights between 0.25m and 0.38m. Gas flow rates ranged from 0.00599m3/s to 0.01312 m3/s. The results of splash measurements showed that splash volume increased with increasing gas flow rate and decreasing bath heights for both tuyere arrangements. However, off-centre configuration resulted in relatively higher amounts of splash. The slag volume in splash (Dph in splash) was also found to increase with decreasing bath heights and increasing gas injection rates. The high speed photography revealed that the frequency and wave amplitude increased with gas injection rate. The experimental results also showed that the frequency and vertical displacement of the rotating wave increased when the tuyere orientation became off-centered. This might be used to explain why off-center configuration resulted in generating higher splash volumes than the center configuration.

Jan 1, 2004

By K C. Runge, E Tabosa, A Jankovic

Flotation recovery in geometallurgical modelling is usually predicted based on relationships developed from batch flotation testing of different ore types. These batch flotation tests are usually performed at a constant set of operating conditions and at a fixed feed grind size P80. Usually the laboratory feed grind P80 is that which is expected to be produced in the full-scale flotation circuit for that particular ore type.Flotation recovery will be a strong function of the valuable mineral particle size distribution, with lower recovery of ultra-fines due to poor flotation kinetics, optimum recovery for the intermediate sized particles and lower recovery for the coarser particles due to poorer liberation of this fraction. This valuable mineral particle size distribution is not characterised by the solids P80 parameter alone. The proportion of material in each class will also be a function of the slope of the size distribution and the degree of preferential grinding of the valuable mineral in comparison to the total mass of the ore.This paper will demonstrate (through example) how these other size distribution parameters can change with a change in grinding technology, the size and operating conditions used in a grinding unit and the characteristics of the classification technology used in conjunction with the grinding unit. Case studies will be presented that demonstrate how this can result in these parameters being very different in laboratory testing to that produced in the full-scale comminution circuit. A rule of thumb that laboratory rod milling rather than ball milling produces a size distribution most similar to that of the full-scale grinding circuit will be challenged. These differences can result in a significant change in flotation recovery. It is therefore important to consider these parameters when developing a laboratory geometallurgical program, to enable more accurate prediction of the flotation recoveries that will be achieved in the full-scale process.CITATION:Runge, K C, Tabosa, E and Jankovic, A, 2013. Particle size distribution effects that should be considered when performing flotation geometallurgical testing, in Proceedings The Second AusIMM International Geometallurgy Conference (GeoMet) 2013 , pp 335-344 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 29, 2013

The profitability of gold placer mining operations is determined by more factors than many placer miners recognise until too late. This study is based upon field trips to examine the efficiency of mining and transportation of gravels to a wash plant by bulldozer and truck. At the examined open pit mining operations gravels were washed using water canons, derockers or scrubber trommels. The objective of this study was to determine the relationship between unit costs and power requirements for gravel mining and processing. The author conducted a number of time and motion studies at four open pit gold mines, including the Maly Patom Creek, the Nizhny Patom, the Taimendra Creek and the Kadali-Makit in the Bodaibo District, Irkutsk Region and collected and processed field data. As a part of the study significant correlations were determined and empiric regression equations derived between: power requirements (Ep) and unit cost (Cp) for gravel processing; and total power requirements (Et) and total unit cost (Ct) for gravel mining, transportation and processing. Based on the results of the study it was proposed to modify the current mining procedure methods and to replace water cannons at the Maly Patom Creek with a PPM-5 derocker. Introduction of the proposed technology resulted in 2004 in US$385 000 mining cost reduction at the Maly Patom Creek over a short-time Siberian mining season.

Jan 1, 2007

By T. Mills, M. Evatz

Newmont’s Lone Tree Mine, located in northern Nevada, is an open pit gold operation supported by a standard truck and shovel mine production fleet (Caterpillar 785Bs and Hitachi EXs). Newmont continues to search for ways to decrease capital expenditure and lower operating cost, hence the implementation of the Ultra Light Body (ULW) concept. The fundamental idea associated with the ULWs is the increased payload potential. Economic justification comes from successful achievement of additional payload, within the truck’s rated GVW, without changing its loading and hauling characteristics. The increase in payload capacity allows equivalent tonnage movement using fewer trucks.

Jan 1, 2001

By Jafar Safarian, Manish Kumar Kar, Ahmad Hassanzadeh, Casper van der Eijk

To minimize the carbon footprint in the industrial valourization of bauxite residue, hydrogen was used as a reducing agent. The current study experimentally investigated hydrogen reduction of bauxite residue-CaO sintered and self-hardened pellets at 1000 °C, along with magnetic separation of these reduced pellets for iron recovery. Calcium was introduced to bauxite residue to form leachable calcium aluminate phases with the existing alumina in bauxite residue. This involved the addition of either CaCO3 or a mixture of CaO and CaCO3 while maintaining the fixed Ca content during pelletization. The former underwent sintering at 1150 °C, while the latter was self-hardened through the cementing effect of CaO in exposure to moisture and air. Both types of pellets were reduced in a thermogravimetry furnace at an elevated temperature under similar conditions. The pellets were characterized by the X-ray diffraction (XRD) method and scanning electron microscope (SEM) coupled with energy dispersive spectroscopy (EDS), and their physical and mechanical properties were measured via standard techniques. During hydrogen reduction, a negligible amount of gehlenite ( Ca2Al2SiO7) was formed in the self-hardened pellets, while this phase dominated in the sintered pellets. Alumina in the bauxite residue converted to mayenite phases during reduction in both the pellet types; however, reduced self-hardened pellets had a higher amount of alumina containing mayenite leachable phase. The two pellets showed similar reduction behaviour, while different chemical, physical, and mechanical properties were observed. The magnetic properties of milled reduced pellets were examined through a Davis Tube magnetic separator in a wet environment under a constant magnetic field of 800 G. Higher iron recovery was observed for the self-hardened reduced (41%) pellets than for the sintered pellets (27%).

Nov 23, 2023

By Gene Mattila, Ron Boyd

Introduction Rock bolting is a general term that encompasses pinning, roof bolting, rib bolting, and cable bolting. Rock bolting contributes the highest nonproductive cost and is the most significant safety risk in mining today. Total installed bolt costs are now more than $20 per bolt. Rebolting or repair work doubles this. And bolting accidents account for about 30% of all fatal and nonfatal lost time accidents. It is only logical, then, for the mine operator to install the highest quality anchors possible. Every bolt on the market - tensioned, untensioned, full contact, or point anchor - has designed capabilities in axial, shear, and fluctuating ground conditions. Bolts will provide a designed support if material specifications and installation procedures fall within the manufacturer's recommendations. This also assumes that there are no unexpected geological or operational changes. Proper material specifications and installation procedures should guarantee the mine operator a quality anchor. To help assure this, the Mine Safety and Health Administration (MSHA), the US Bureau of Mines (USBM), and the AMC Manufacturers group developed stringent ASTM materials standards. These govern mechanical point anchor bolts and bolts using threaded and nonthreaded rebar. There are also minimum standards that suppliers must comply with in the manufacture of frictional bolts. However, most frictional bolt standards are based on the manufacturer's own specifications. It would be impossible to set uniform regulations on frictional bolts. Each bolt provides support in a slightly different manner and thereby requires different material needs. Quality Control In any case, material quality can generally be assured through tight order specifications and periodic quality control checks. The problem with installing quality anchors does not solely lie with the material. It includes all the events that occur from the time a bolt leaves a manufacturer to the time it is installed. The anchoring ability of any major type bolt can be affected by transportation, storage, mechanical performance of installation equipment, installation demands of the operator, proper drilling accessories, and local geology. Table 1 and Fig. 1 document the negative events that can affect each bolt and the degree to which it can be affected. Drilling a hole of oversized diameter, for example, is an operator controlled factor. It will adversely affect most bolts. An over-sized diameter reduces, or even eliminates, the anchorage capability of a mechanical shell. It will reduce the support capacity of a split set to nearly nothing. And it will cause a partial or incomplete grouting around the resin bolt. Frequently, due to bit wear, holes are drilled undersized in diameter and may not be properly cleaned out. In this case, it will be difficult for the operator to push the mechanical shell into the hole, thereby causing him to drill another hole. In cases where the bolt is obviously not providing anchorage, operators will probably drill another hole and place another bolt. Rock Bolt Regulations It is the situations where a bolt is improperly installed in the hole, but gives the appearance of a good bolt, that are frightening. In fact, some safety officials and engineers believe that only 50% to 60% of installed bolts are working. These bolts cause irregular ground load transfers in other bolts and overload them. This causes hazardous situations and overstressing. It is these bolts that fail prematurely 200 to 255 mm (8 to 10 in.) behind the face and possibly contribute to the high accident rate. The government and several manufacturers recognized the need to check the integrity of rock bolts. These programs are in different stages today. Tight torque regulations have been established for mechanical

Jan 4, 1985