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By M. Alley, P. Schroeder, L. Zelazny, S. Nagle

The Old Hickory mineral sands ore body in Virginia (>2000 ha) currently supports a productive rowcrop agriculture. Reclamation experiments were implemented on pilot-scale mining pits from 1995 to 1998. Half of each mining pit was topsoiled (25 cm) while the remaining half was left as either (1) mixed tails/slimes or (2) re-graded subsoil over tails/slimes treated with 112 Mg/ha yardwaste compost. Post-mining productivity was estimated to be reduced by 23%, 3%, 27%, and 20% for each crop (wheat/soybeans/corn/ cotton) in sequence when compared to undisturbed soils. Variability in surface texture and compaction limit restoration of soil productivity.

Jan 1, 2002

By A Proudman, J Isles, R Lawolo, G Forster

PT Nusa Halmahera Minerals (PT NHM) is an Indonesian joint venture company owned by Newcrest Mining Limited (75 per cent) and PT Aneka Tambang (Persero) (25 per cent). PT NHM operates two underground gold mines, Kencana and Toguraci at the Gosowong mine site in North Halmahera District, North Maluku Province, Indonesia.Due to several major stope overbreak events at Toguraci, the mining method and sequence was reviewed. The method was changed from a bottom up Avoca stoping approach using waste rock as fill material and leaving sill pillars, to a top down open stope method using paste fill without sill pillars.The revised mine plan originally had stoping progressing from the current level to the base of reserves followed by extraction of the sill pillars in the upper levels at the end of mine life. Geotechnical modelling was conducted on the revised mine plan and it was recommended that the sill pillars be taken early due to the risk of them overstressing and being difficult to recover if left until the end of mine life. The backfill material above the sill pillars is unconsolidated run-of-mine development waste rock or unconsolidated rock fill (URF).Due to earlier attempts to extract sill pillars under URF not being successful in the upper areas of the mine, it was stipulated that the URF above the five remaining pillars must be consolidated prior to stoping. Cost benefit analysis showed this to be justifiable due to the high-grade nature of the ore locked up in the pillars.Grout was initially planned to be injected in one pass through self-drilling anchors (SDA) in a 2.5 m pattern. Initial attempts to do this failed due to the SDA aperture and grout consistency. A modified method was employed using poly pipe installed into open holes that were drilled through the sill pillar and into the URF from below. Repeated application at each location ensured full grout consolidation. Drill holes were used to confirm grout thickness as well as add more grout. Design thickness was based on stress analysis.Despite different URF types being present due to varying quantities of fines, the process was similar for each area. The presence of paste in the URF of one target affected the quantity of grout it received.With a design thickness of 4–5 m of consolidation in the URF, depending on stope width and dip, each consolidated pillar (to date) maintained integrity during the stope operation. Some fall of material was anticipated; however, quality assurance / quality control (QA/QC) during grouting, staged stope extraction and careful blast design minimised fall of material.CITATION:Proudman, A, Isles, J, Lawolo, R and Forster, G, 2017. Rock fill consolidation for the extraction of sill pillar stopes at Toguraci gold mine, in Proceedings 13th AusIMM Underground Operators’ Conference 2017, pp 199–210 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 16, 2017

By D. C. Bailey, D. I. Hodson

Paper presented at the annual joint meeting of the IMM and IMinE held in Harrogate, 31 January 1991. The Neves-Corvo copper-tin mine is the highest-grade large-tonnage copper operation in the world and is now arguably the most significant producing base-metal mine in Europe, rating in the top ten world producers of both copper and tin. A description is given of the initial exploration work and the development of the project from the orebody discovery in the Iberian Pyrite Belt of southern Portugal in 1977 to full operating capacity in 1990. The chronological account includes discussion of the main organisational, financial, technical and environmental decisions and of the various difficulties encountered, and the challenges of the future are identified. The project was brought in on schedule and within budget

Jan 1, 1991

By Gungor Tuncer, Ali Kahriman, Abdulkadir Karadogan, Umit Ozer, Ertugrul Kaya, Kagan Ozdemir

The results of bench blasting on quarries, have important influences on the performance of subsequent mining activities such as loading, hauling, crushing, grinding etc., and the total costs of that mine. Drilling and blasting is an important step in mine producing chain and its results such as fragmentation, muckpile shape and looseness, dilution, damage and rock softening effect the efficiency of downstream processes. Within the scope of this study, the three different hydraulic excavators’ working performances have been studied considering the particle size distribution after bench lasting at a quarry in Istanbul Turkey. Five shots have been realized on the quarry and average particle size of muckpile after each shot were assessed by digital image analysis method. Bucket loading times have been measured by stopwatch study after five shots. Excavator bucket loading times were also estimated by bucket loading time estimation formula which was published in ISEE 2007 Conference, Ozdemir et al. 2007. Estimated and measured times have been statistically analyzed and the accuracy rate of formula has been determined for five events.

Jan 1, 2008

By Neil J. Mansfield

Tool manufacturers are required to provide declarations of vibration emission values in order to sell their tools within Europe. To ensure that users can compare results obtained from different manufacturers, the declared values must be obtained using a methodology as specified in the relevant test code (such as in the ISO 8662 series of standards). In most cases, the vibration emission values obtained using test codes under-estimate the vibration that an operator will be exposed to when using the tool on a work-site. A further problem with manufacturers’ data is that usually only a single value is provided for a tool. This is despite many factors affecting the vibration emission, including inserted tool type, work piece, operator technique, tool condition. New improved test codes are in the process of being developed. In order to provide guidance to users on how to interpret manufacturers’ data, a Draft CEN Technical Report (Draft CEN/TR 15350 (2005)) was developed. Part of the CEN/TR provides multiplication factors for combinations of task and tool type. For example, data obtained from electrical hammer drills (tested according to EN 60745-2-6:2003) should be multiplied by 2, for hammering applications, in order to obtain an estimate of the vibration emission during work. In response to concerns from industry, the UK trade association OPERC have, in collaboration with hire companies and tool manufacturers, established a freely accessible online database of tool emission values based on independent tests carried out under simulated work-site conditions. This paper reports some of the data obtained from electrical hammer drills, highlighting the range of emission values that can be obtained for a tool. Data from many other tool types are also included in the database. Methods Tri-axial hand-arm vibration was measured at both handles of each of 19 electrical hammer drills, in accordance with ISO 5349-1 (2001). Each tool was measured with three experienced operators and at least 5 runs were completed for each operator. Tools were tested using a range of appropriate new bits from 4 to 40 mm diameter. The minimum number of bits for a tool was 3; the maximum number was 29. 146 tool / bit combinations are reported here, representing about 2200 individual 6-axis measurements. Operators were required to drill vertically into a concrete block with a compressive strength after 28 days of 40 N/mm². Two drills (1 and 15) were battery powered; others were powered using a 110V transformer supply.

Jan 6, 2006

By Mark William Withey

"Blasting safety in surface mining is a necessary industrial goal. Proper techniques and disciplinesexercised during the blasting process enhance the probability of achieving that goal.The blasting process is generally a series of repetitive steps performed by all blasting technicians.Because habits formed during this process influence safety, production and cost, proper habitsbecome compulsory."

Jan 1, 1995

By D Craw, R Norris

The Golden Bar pit is part of the Macraes mine, the largest gold mine in New Zealand. The Macraes mine is mainly developed in the late metamorphic Hyde-Macraes Shear Zone. However, the Golden Bar pit was developed in a mineralised zone about 200 m structurally above the Hyde-Macraes Shear Zone. The host rocks are dominated by massive quartzofeldspathic schist (5 - 10 m layers) with minor intercalations (1 - 5 m scale) of micaceous schist. All the host schist has disseminated pyrite (micron to mm scale) from pervasive late metamorphic hydrothermal alteration.   In addition, there has been more localised (metre scale) relatively intense hydrothermal alteration controlled by shears subparallel to foliation. Gold occurs in three types of mineralised rock: silicified schist, which dominates, fissile black sheared rock, and cross-cutting quartz veins. There are three principal structural zones in the deposit: 1.   a strongly sheared micaceous and graphitic zone at the base of the structural sequence; 2.   a weakly deformed schist zone cut by relatively flat-lying (= 20¦ dipping) quartz veins; and 3.   a stack of imbricated schist pods and deformed silicified schist confined between two sub-parallel low-angle (ca. 30¦) shears.   These structural zones were juxtaposed during thrust deformation. The lower black shear zone has m-scale folding and over-thrusting packages, kink folding and foliation stacking with associated striations that imply west-directed thrusting. Striations and imbrication geometry between the upper two shears imply oblique thrusting that was directed towards the west. Post-mineralisation normal faults locally disrupt the mineralised zone. The Golden Bar pit mineralised zone has some of the structural and mineralogical characteristics of the main Hyde-Macraes Shear Zone, and may be genetically related.

Jan 1, 2006

By Thomas O. Meyer

Laboratory test results are presented for the determination of elastic modulus made with the NX borehole jack. Determinations were made in massive pieces of aluminum so as to satisfy conditions of linear elasticity, homogeneity, and isotropy. Results indicate a large discrepancy (approximately 200 percent) betwen material modulus as determined by laboratory block-loading tests and that indicated by the jack under relatively ideal circumstances. Effects due to interface friction, finite block size, incomplete plate-bore contact, and other factors were examined and found not to be significant. Finite element results indicate that assumed loading conditions for the jack differ from those actually occurring, at least in materials of high elastic modulus (approximately 10 million psi), such as quartzite.

Jan 1, 1974

By A. Iannacchione

Longwall mining began within the Pittsburgh coalbed of Pennsylvania (PA) in the early 1970?s. Originally, longwall mine layouts were designed to fit within existing room-and-pillar operations, where panels were limited to that of typical production panels, e.g. a few hundred feet wide by a few thousand feet long. In these early designs, the gate entries were comprised of chain pillars similar to those found along main entry systems or within pillar recovery sections. The modern longwall mine has little in common with their 1970?s predecessors. Today the entire mine layout is designed around the longwall mining method. The longwall panels have grown to well over one thousand feet in width with lengths now measured in miles. In addition, development pillar sizes and shapes are matched to specific performance requirements.

Feb 27, 2013

By C. K. Kocsis

A study was undertaken to evaluate the thermal environment and determine the contributing heat sources within the existing 153 Orebody of Coleman/McCreedy East Mine which would then enable the prediction of the climatic conditions in the mine?s future 170 Orebody. The study entailed continuously monitoring, for a 2-week period, the dry-bulb temperature, humidity and barometric pressure along the existing intake airways from surface down to the 4810 Level and throughout an active mining area on that level. Furthermore, the mining activities and associated equipment that could be responsible for any changes in the temperature and humidity conditions within this mining area were also monitored. These data were then used to develop and validate a climatic simulation of the mine?s mechanized cut and fill mining method. This climatic model was then transposed to the future 5700 Level in the deeper 170 Orebody, to determine the environmental conditions within its production workings and along the main haulage route ascending from that area. Based upon the simulations performed it was shown that the level of ventilation planned for the new orebody should be able to maintain adequate environmental conditions. However, this study did emphasize the importance of the appropriate distribution of airflow in workplaces that used auxiliary ventilation in order to sustain a suitable thermal environment.

Jan 1, 2009

By J. Geyer, B. T. Burton, H. W. Newcomen

For open pits developed in tropical climates the upper portions of the pit walls are often located in residual soils and weathered rock. Due to the presence of relict structures, and the relatively low strength of the residual soils and weathered rock, design slope angles in these materials have to be developed by blending the results of kinematic assessments of geologic structures with rock mass stability analyses and traditional soil mechanics. Background information on the engineering considerations of residual soils are discussed, and the results of geotechnical assessments for proposed pit design of the Brisas del Cuyuni Project, located in southeastern Venezuela, are presented.

Jan 1, 2000

By Susumu Uchiyama

This machine was manufactured for excavation of the Roppongi station of the Tokyo Metro Line 12. This machine can excavate both the platform and the rail track at the same time in locations where the traffic is congested and has following features: A multi-centered TBM of the size of 7.06(H) x 13.18(W) m The world first H&V TBM, with right and left machine parts articulated vertically and separately for position control A large erector for simultaneous erection of column and central segments of special shape to construct a highly reliable tunnel This paper reports the design concept, construction, and work record.

Jan 1, 1999

The management of risks associated with slope instability is an essential process in the safe and economic operation of open cut mines. The ?slope stability radar? (SSR) has been developed to better manage those risks. The SSR remotely scans rock slopes to continuously measure any surface movement and can be used to detect and alert users of wall movements with sub-millimeter precision. The high level of movement precision and broad area coverage of the rock face can allow for a better understanding of the geomechanics of slope deformation, including magnitude of potential failures and additional warning time of impending instability. Additionally, radar waves adequately penetrate through rain, dust and smoke to give reliable measurements, 24 hours a day. SSR systems have been deployed in many metalliferous and coal mines throughout the world. Over 80 systems are currently being used with current operations in 17 countries. Greater than 200 rock slope collapses and waste dump failures (from several to millions of tones) have been monitored, with ?warning? precursor movements recorded by the SSR. This technology enables a radical change in the management of risks in open cut mining. This paper will present recent case studies of the SSR operation in coal and metalliferous mines. Slope Stability Hazards and Rock Slope Monitoring in Open Pit Mines Surface mining often creates excavation boundaries which are steeper than those would be created under natural processes in the same rock type. They are also undertake quite rapid stress changes during mining and are influenced by significant external forces such as blast vibrations and changes in groundwater. These conditions are likely to allow some degree of instability to develop in the pit slopes. The hazards associated with this instability of pit slopes can be extremely variable in character; ranging from small cobble sized rocks dislodging and falling from mine benches to complete collapse of entire pit walls. It is usually the position of mine geotechnical or geological staff to assess these hazards for mine operations. The assessment of hazards due to unstable slopes can be done by visual observation of precursor signs such as fracturing or rilling of loose material. A more reliable indicator of instability involves the quantitative measurement of outward movement and acceleration of material as an instability mechanism develops. There is strong evidence that small precursor movements of a rock wall occur for an extended period prior to rock slope collapse (Hoek and Bray, 1981). Development of a monitoring system, adopting acceptable slope deformation criteria coupled with warning systems and design of stabilization or risk reduction measures if appropriate has become a standard method of dealing with slope instability. Framework for Slope Hazard Management using the SSR A framework for the management of slope hazards is shown in Figure 1 (discussed in detail in Harries, 2007). Although this framework is specifically for use with the Slope Stability Radar, a similar framework could be developed for other monitoring technologies for monitoring slope deformation. An important point to consider is the sequential nature of the hazard management framework from selecting the context ? identifying hazards ? analysing hazards ? evaluating hazards ? treating hazards. [ ] Once the context of the system has been developed (for example ? to monitor for rapid bench scale collapse), the identification of hazards is the next stage in hazard management. Identification of slope hazards can be relatively straightforward due to the visual nature of the information that the SSR system produces. The areas of excessive wall deformation are highlighted in color maps of the mine slope with a co-aligned photographic image to review the actual site of interest. The responsible personnel must then analyses this data, usually reviewing the deformation style and rate, as well as the size of area that appears to be failing. This information can then be evaluated against organisation protocols (for example it may be only a small sub-bench failure that will be ?caught? by existing benches and hence can be categorized as a lower risk). Once these hazards are evaluated they can then be treated if required. Two major options are available to reduce the risk during this treatment option. We can reduce the likelihood of failure ? for example by buttressing the failure or de-pressurising of water pressures in a slope. However, the most likely plan of action is to reduce the consequence of the failure. One of the most successful opportunities to reduce the consequence of a slope failure is to ensure that no personnel or mining equipment is present at the location of a slope failure when collapse finally occurs. This has been the major benefit of the SSR system to the different mining operations around the world. SLOPE STABILITY RADAR TECHNOLOGY OUTLINE Slope Stability Radar system specifications and capabilities: ? Measurement precision +/- 0.1mm (std.) ? Perforated dish and other features enable it to operate in high wind conditions ? Measurement range is 1700 metres ? Scans 270 degrees horizontally and 120 degrees vertically ? 1 ? 30 minute repeat rate for area scans (varies according to area being scanned ? RF exposure (<0.05mW/cm2) is a factor of 200 below the acceptable limit for the general population (1 mW/cm2) ? Tolerant to vibration and mining equipment ? The system weighs approximately 1500 Kg ? Remote area power supply enabling unmanned, continuous 24/7 monitoring ? Real time monitoring

Jan 1, 2009

By E. Morrice

As part of its efforts to maximize process efficiency and decrease energy requirements in mineral processes, the Federal Bureau of Mines has investigated methods for beneficiating bastnäsite ores from Mountain Pass, Calif. The results of a bench-scale study on the flotation of bastnäsite ore are presented. Flotation at a pulp temperature of 40° to 45°C yielded a cleaner concentrate containing 58.7 pct RE203 at a rare-earth recovery of 67 pct. These results are comparable to those reported for the commercial beneficiation method in which the pulp is heated to boiling before flotation. An attempt was also made to recover a barite flotation concentrate from the bastnäsite rougher tailing. Concentrates containing only 70 pct barite were obtained.

Jan 1, 1982

By M. Fustic, A. H. Paul

The geology of the Muskeg River Oil Sands Mine is described to demonstrate how geologic modeling aided in resource evaluation and material characterization during the feasibility and construction stages. Bitumen is hosted in the clastic sediments of the Cretaceous McMurray Formation which is informally divided into members representing the fluvial, estuarine and marine environments of this transgressive sequence. Lithofacies mapping of drill core provides further detail to classify materials as resource type, for extraction suitability and for construction suitability. Geologic interpretation provides tools for the engineers to produce effective mine plans and for the operators to execute them during mine development.

Jan 1, 2002

By Sami Hindstrom, Travis Orser, John Starkey

"A great deal of effort has been expended in the last 11 years by the authors and others to identify the correct power required to grind ore in a SAG mill, by developing ore hardness variability functions for the ore body to be mined. This paper describes the next step in the SAG design process, which is choosing a SAG mill to accomplish the required specified design performance.At the test work stage, each ore hardness data point is a unique and discreet bit of information that represents the energy required to grind the ore represented by that sample at optimum SAG conditions. When the design criteria are set, it is now considered necessary that the required power draw in the mill chosen be provided at optimum conditions as well.Based on actual 3-ft diameter pilot plant results, we can now show that the grinding efficiency and throughput drop when the optimum load limit is exceeded. Future work is expected to show a similar reduction in efficiency when exceeding optimum speed. Providing specified power by exceeding these optimum conditions is not adequate and leads to a lower cost but underperforming SAG mill.It is possible in some cases to modify a SAG circuit to force it to perform better, but this often takes time, much effort and costly modifications. In most cases the capital spent on a properly sized SAG mill may prevent production losses that would be many times greater over the extended start up period. This is the direct result of choosing the proper SAG mill at the start.INTRODUCTIONThe sequential steps required for the design of any grinding circuit, have changed dramatically in the last seven years. Previously, it was considered necessary to do six foot diameter SAG mill pilot plant tests on representative samples of minus six inch ore (F80), in order to understand design requirements. It can now be shown that this method is costly and not likely to produce the desired result because it is difficult to recover six-inch material from deep within a new mine. The representative nature of the sample is often compromised when the sample is selected."

Jan 1, 2003

By C. A. Pickles, J. M. Toguri

The refining of two molten Ag-Cu alloys, containing 7.5 and 11.2 mass% Cu, with a borosilicate slag was investigated using both air and pure O2 to determine the effects of the following variables: flow rate of oxidising gas; injection time; mass of sodium borosilicate flux; and injection of the oxidising gas into either the slag or the metal. The results demonstrate that more than 90% of the Cu can be removed with Ag losses of less than 5% and that refined metal with a Cu content of less than 0.5 mass% can be produced. Ag losses to the slag increased with the amount of Cu removed, and a mechanism is proposed to account for this

Dec 1, 1995

By C. Ozgen Karacan, William P. Diamond

In This Chapter [Measuring the gas content of coal Predicting gas emissions based on geologic and coal reservoir property data Determining the gas storage capacity of coalbeds and other gas-bearing strata Methane drainage borehole monitoring to forecast the remaining gas-in-place and the influence on mine emissions Forecasting gas emissions during mining as a function of mining parameters and Gas emission prediction based on numerical simulation] This chapter provides guidelines for determining the gas content of coalbeds, estimating the gas-in-place, and predicting gas flow and emissions before and during coal mining operations. The techniques are discussed briefly in the following sections. However, detailed information on the techniques is provided in the cited references. INTRODUCTION Coalbed methane, if not properly controlled in the underground mine environment, is a safety concern due to the potential risk for an explosion. This is a particular problem during longwall mining, where the high rate and volume of coal extraction can result in the release of large amounts of methane from the mined coalbed and other adjacent gas-bearing strata. The variability and potential hazards of these sometimes unexpectedly high gas flows provide the impetus to develop methods to predict methane emissions into the underground workplace. A forecast of the volume of gas that might be released during coal mining is helpful for designing ventilation systems and for implementing optimum methane drainage strategies to help mitigate expected gas emission problems. A complete assessment of the need for methane drainage prior to mine development generally requires both an empirical and a theoretical approach. If there are active mines in the general area with similar geologic conditions and coal characteristics, a review of gas problems in those mines provides an initial insight into the level of gas emissions to be expected at a new location. In addition, relatively simple methods exist to determine the in situ gas content (volume of gas per unit weight of coal) of the coalbeds in a particular mining area, as well as the gas-in-place (volume of gas in the coalbed(s) within a defined geographic area).

Jan 6, 2006

By S. G. Shabestari, S. Saeidinia, P. Mokarian

"The experiments were conducted to investigate the influence of various process parameters in rotary degassing of aluminum alloy AS5U3G. Degassing was carried out by use of argon inert gas. The flow rate of the gas was investigated in the range of 12 to 21 L/min., 350 to 500 rpm rotor speed, 6 to 12 minutes degassing time, and 755°C to 780°C melt temperature. Degassing efficiency was evaluated through measuring the density of RPT samples. By optimizing the process parameters, RPT sample density was increased up to 2.73 g/cm3, which guaranties the production of high-quality aluminum castings. Tensile strength and elongation were increased considerably by degassing process. Scanning electron microscopy (SEM) shows small size and well-distributed microporosity. IntroductionOne of the major problems associated with the casting of aluminum alloys is the formation of gas porosity. Aluminum and its alloys are very susceptible to hydrogen absorption. Because solubility is much lower in solid state than liquid state, hydrogen atoms leave their position during solidification and combine together to form hydrogen molecules. These micrometer- scale cavities, called microporosity, cause the reduction of mechanical properties, particularly strength, ductility, fracture toughness, fatigue resistance, pressure tightness, and corrosion resistance (Laslaz, 1991; Jorstad and Rasmussen, 1993; Sigworth and Wang, 1994; Korasani, 1996; Anson and Gruzleski, 1999; Zhang et al., 2000). To avoid these problems and produce good quality castings, it is necessary to remove hydrogen from the melt. Several methods are available for degassing aluminum melt. One of them is a rotary degassing system using inert gas, e.g. Ar or N2. Various thermodynamic and kinetic factors govern the removal of hydrogen from molten aluminum (Anderson, 1987; Sigworth, 1987; Pignault, 1993; King and Reynolds, 1995). These are some parameters that affect degassing efficiency. Some of them are discussed here and some are examined in this project."

Jan 1, 2004

By T. G. Joseph

Mines must manage variability from within their operations as well as those imposed upon them by external forces. This paper provides a framework to manage both internal and external variability through the review and adjustment of the mine plan where applicable to minimize lost production and maintain operational safety.

Nov 1, 2011