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By P. C. Hodges

WHILE there is very little romance connected with the operation of a stone quarry, yet to those who have participated in the growth and development of a business that has been a pioneer in its field and that FIG. 1.-LOCATION OF PROPERTY OF MARBLE CLIFF QUARRIES COMPANY. has been in continual operation two-thirds of a century there is bound to be a background of experience, which if properly capitalized, brings progress as well as satisfaction in its accomplishment.

Jan 1, 1936

By Jean-Michel Rendu

On April 30, 2003, representatives of the Society for Mining, Metallurgy and Exploration, Inc.(SME) met with the staff of the Division of Corporation Finance of the U.S. Securities and Exchange Commission (SEC). The meeting was held to discuss issues related to the estimation and publication of mineral resources and ore reserves. It took place in the Washington, D.C. offices of the SEC. SEC staff members included Roger Schwall, assistant director, Division of Corporation Finance; Jill Davis, senior accountant; Roger Baer, senior mining engineer; Carl Chavez, chief mining engineer; Ron Winfrey, senior petroleum engineer; and John Murphy, senior petroleum engineer. SME committee members included JM Rendu, independent consultant and chairman of the SME Committee on Resources and Reserves; Ian Douglas, manager of deposit modeling and reserve estimation with Newmont Mining; Art Graham, director and mining engineer with Deutsche Bank Trust Company Americas; Larry Smith, chief geologist with AMEC Mining and Metals Consulting; and George MacDonald, vice president of exploration with Freeport-McMoRan Copper and Gold.

Jan 1, 2003

For its Hellyer project, Aberfoyle sought to establish a high degree of confidence in all aspects of the feasibility study. Accurate capital and operating cost estimates were an integral and essential part of the study. They were crucial to the evaluation of investment returns and competitive cost position. Both the capital and operating cost estimates were based on the comprehensive data available from exploratory mine development, pilot plant operation, detailed engineering design, and Aberfoyle's experience in the area. One of the principal project objectives was to maximise operating productivity and efficiencies and this created capital cost alternatives which needed thorough cost-benefit assessment. The estimated capital cost for construction and development was $90 million, and the project was completed virtually on time for $80 million. The first year's experience indicated that operating costs would settle to well within 10% of the estimates.

Jan 1, 1990

A production drill rig at a site in Sweden was equipped with a monitoring system to estimate production statistics. Drill running times, drill hole depths, penetration rate, rotation speed, thrust, air percussive pressure and torque pressure were recorded for every 10 mm of hole length for a period of 5 months. Operational activities (rock penetration, tube handling, moving and set-up time) accounted for almost 62% of the time consumed, the remaining 38% being different types of stoppage. Drilling itself, as opposed to handling, took less than 24% of the total time. The results give an indication of how the overall drilling capacity can be increased and which activity should be emphasised. The impacts of automation, new developments and reorganisation on production can also be predicted. Average production was found to be around 100 m/day, but varied between 50 and 160 m/d. The company's specification for the drill rig, before purchase, was that an average production rate of 130 m/d be attained over two consecutive months of drilling - a rate which monitoring shows will not be achieved

Jun 19, 1905

By Brian T. Burch, Paul N. Worsey

When fired, submerged Linear Shaped Charges (LSCs) lose all effectiveness in the cutting of steel. Users in underwater applications have reported having to substantially increase the charge size to the point where brute explosive force shears rather than cuts the target.

Jan 1, 2015

To provide a flexible, low-maintenance, low-cost continuous mine air monitoring system as an alternative to conventional fixed-point, electronic monitoring systems. Background Continuous monitoring of mine air is highly desirable (and often required) not only to improve miner safety, but also to increase mine productivity. Continuous air monitoring systems can provide earl" warning of developing fires, excessive methane accumulation, reduced oxygen levels, and other developing problems in remote, inaccessible or sealed areas, and in active workings. Fixed-point, electronic systems require one Sensor for every gas to be measured at every defined monitoring location. As the number of monitoring locations increases, so does the cost, along with increases in the time required for routine maintenance and calibration of the sensor System cost also increases when monitoring locations are identified for mine areas requiring permissible electrical equipment.

Jan 1, 1988

By F Kelly

"A central composite experimental design was used to determine empirical relationships between the operating variables of a ball mill and the characteristics of the ground products. The independent variables selected for the study were the work index and size distribution of the feed material, the feed rate and the pulp density. The measured responses included product size distribution, product surface area, viscosity of pulp discharge, weight of dry solids in mill, mill time constant and emitted sonic vibrations. The ball mill was operated continuously in open circuit during each experimental run.Multivariable regression techniques were used to derive linear empirical models from the experimental results. Of the four factors studied, the grindability of the feed material significantly affected every measured response. The size distribution of the ground product was affected significantly by the feed rate and the size distribution of the feed, but not by the pulp• density. Feed throughput capacity and pulp density were significant to the viscosity of the pulp discharge, the weight of solids in the mill and the mill time constant. Significant first-and second-order models are given for the above relationships. The relation-ships between the characteristics of the sonic vibrations of the operating mill and the various dependent variables have not as yet been established."

Jan 1, 1970

By Ing. R. Geutskens

Zinc-bearing dust from blast furnace gas-scrubbing operation is a well known environmental health problem in the iron and steel industry. Hoogovens IJmuiden (The Netherlands) developed processes to solve this problem. 90% of the zinc is concentrated in 25% of the dust (the fine particle fraction) by means of slurry-hydrocyclonage. The resulting fine particle fraction is treated in a pressure leaching process. Hydrochloric waste pickle liquors have been used as a leaching liquor. By creating optimal leaching conditions more than 98% of the zinc is brought into an acid' solution, whereas nearly all the iron remains in the residue. The zinc can be separated selectively from the solution. The residue can be reused in the blast furnace (via the sintering plant). In the article information is given about the research work and -results of the pressure leaching process as well as the possibilities of the process for solving comparible problems.

Jan 1, 1990

By S Williams, N McBride

The aim of these analyses was to utilise the latest available production information for the Wandoo operation to determine the optimal blasting parameters, costing information and predicted results (fragmentation, muckpile shape and vibration). Much of the data used as input was made available from the information gathered during the excavation of the test pits during 1995. As with any æplanning stageÆ analyses, many assumptions were required to fill gaps in the available data. The calibration of computer blasting models was conducted using data from the last of the test pit blasts (this blast was carried out in material that closely resembled rock expected to constitute the bulk of the base case design pit). This calibration was refined using observation and experience based judgements to extend the accuracy of the predictions as far as possible. All the simulations were performed assuming that the blastholes will be wet (as instructed by BGM). This has consequently determined that the appropriate explosive for the simulations and the actual blasting will be a pumped emulsion/ANFO blend. The following blast design parameters were derived by determining a likely range of powder factor (from other similar scale gold mines) with a median value of 0.62 kg/m 3 being the powder factor arrived at through the field optimisation process conducted during the test pit excavation. The total swell/displacement percentage (volume after blast/volume before blast) largely determines the diggability of a well fragmented muckpile. Consequently where fragmentation is good any reduction of total swell will probably reduce diggabilty. Any reduction of diggability can and should be traded off against increased gold recovery. In most gold mines, diggabilty will be substantially compromised at the balance which produces the most profit. When these considerations are combined with the cost to drill and blast material, processing costs and the achievement of æacceptableÆ diggibility, the analysis becomes extremely complex and qualitative. Many trends are evident including:the relationship between powder factor (for each blast design) and maximum forward displacement (throw), see Figure 1; and the relationship between powder factor (for each blast design) and drill and blast cost per BCM, see Figure 2. The following blast designs are matched to each of three ore-tonnage alternatives (as supplied by BGM) (Table 1). They are all based on powder factor of 0.62 kg/BCM and are considered optimum from a mining æpoint of viewÆ. The costing of each design includes drilling, explosives and labour cost to load, stem and tie-in holes. Using the available rock data and relying heavily on the wall exposures in the test pit(s), it is the opinion of the author that between 20 per cent to 40 per cent of the final walls excavated in the Wandoo pit would be likely to have the angle between the dominant foliation and the wall plane that is well suited to presplitting and be cost-effectively served by presplit technique. The drilling cost for 125 mm blasthole is about $11.40 per drilled metre and the cost of 40 mm presplit explosive is about $140 per 25 kg (17 metre) case. This implies that the cost to drill and blast presplit on the above parameters would be about $10.90 per square metre of presplit face. The remaining wall area would be most cost-effectively delineated using modified production blasting techniques. This requires the use of modified parameters such as smaller hole diameters, smaller spacing/ burden, higher powder factor, free face firing, less rows per blast and longer delays. These altered blast parameters would usually be applied uniformly across the entire blast fired adjacent to the wall and this ætrimÆ blast would relieve burden away from the wall, into the pit. In simplest terms, this method relies upon maximised burden relief to minimise blast energy applied to the wall. It is only applicable where the material to be blasted is waste since the throw achieved is inconsistent with ore dilution control. The largest charge weight per hole for these analyses is 360 kg and the nearest domestic structure to the proposed Wandoo pit is about 6100 metres distant. From these parameters and assuming a worst case example of rock controlled site factor, the predicted ground vibration is 0.5 mm/sec that is about five per cent of the lower limit set down in AS 2187 part two (1993). This suggests that for the current location of neighbours, ground vibration will not be an issue for the Wandoo operation.

Jan 1, 1999

Discussion of the paper of Edward Keller, presented at the Butte meeting, August, 1913, and printed in Bulletin No. 80, August, 1913, pp. 2093 to 2115. GEORGE L. HEATH, Hubbell, Mich.:-In the admirable paper quoted, is mentioned the date when the 5-g. assay was first proposed is a standard for refined metal. The author omitted, however, to mention a later paper,1 which contains a proposal for the use of a single " stock solution " of mixed acids, which has been adopted by the American Brass Co. and others. The suggestion is made that such a scheme would save time and manipulation with converter metal. Unless the low current of 0.5 ampere and the large excess of nitric acid, recommended by Mr. Keller, have been proved necessary to hold up antimony-and selenium, the use of the mixed solution which we have adopted will permit the use of 1 ampere of current per square decimeter of cathode surface, until the solutions have become colorless. At this point, we wash the covers and reduce to 0.5 ampere to prevent oxidation or contamination. The ordinary split electrodes have been used. With strong current, at least, the use of an excess of nitric acid tends to the neutralization of the solution before the end of the period, from the formation of too much ammonia., and this is detrimental. Antimony deposits more easily than arsenic, and tends to drag the latter dorm with it. There must, be some such explanation for the fact. that we have made hundreds of deposits of pure cathodes from Lake material containing no antimony, but as. high as 0.5 per cent. of arsenic. By placing an assay beaker in the Frary solenoid, or rotary device, a cathode deposit with less than 0.01 per cent. of impurity is easily obtained in one operation from 3 g. of whitneyite, carrying from 7 to 10 per cent. of arsenic. In such a case, it is necessary to test several times for end point and remove the plate as soon as no brown tint appears (at once) when hydrogen sulphide water is added to the test portion on the spot plate. Edward Keller's proportion,, of solvent figure out as 14 cc. of strong nitric to 11.4 cc. of sulphuric acid (1.84).

Jan 11, 1913

By Shahé Shnorhokian

Numerical modeling is increasingly being used as a tool to examine complex geomechanical issues related to rockmass behaviour and stability of underground openings. The in-situ stress regime and the evolution of mining-induced stresses during the life of an operation, taking into account the planned stope sequence, are topics that can only be handled through the use of a realistic and calibrated numerical model. This paper presents a case study of a deep underground Canadian mine in northern Ontario. A FLAC3D model was constructed and calibrated based on in-situ stress measurements. In the first stage, the evolution of stress regime was studied and compared with the occurrence of rockburst events during a period of three years. The locations of clusters were examined and compared to the mining sequence during that period. In the second stage, stresses within a diminishing ore pillar were examined based on the planned stope sequence and compared to levels deemed acceptable to the mine. Based on the results, an alternative sequence was arrived at to reduce stress accumulation in the pillar, or induce it to occur at a further distance from a brittle dyke in the vicinity of the orebody.

Aug 1, 2013

By M Sandy, P Bamford

The Challenger Gold deposit was discovered in 1995 by Dominion Mining Ltd in a remote western area of the Gawler Craton in South Australia. Mine development commenced in 2002 with initial production from an open pit supplemented by ore from a very small adjacent auxiliary pit. Development of the underground mine commenced in 2004. The orebody plunge of just less than 30 degrees requires careful attention to mine design. Production from the M1 orebody, with less than 800 ounces per vertical metre (ovm) and very limited æsurgeÆ capacity, requires careful scheduling and implementation to ensure continuity of ore feed for milling. Initial production results from underground stopes included more dilution than anticipated. The narrow, high-grade zones proved to have much more complex geometry than early interpretations had indicated. Greatly improved understanding of the mineralisation controls resulted in important changes being made to the underground mine design and method. Coupled with careful control of development and drilling and blasting practices, this resulted in an immediate reduction in dilution and high recoveries of the resource. The benefits of these changes flowed through in the 2005/06 financial year, resulting in a 150 per cent increase in gold production to 108 080 oz and a sharp reduction in cash operating costs to A$280/oz. This was followed in 2006/07 with production of 108 91 oz of gold at an operating cost of $309/oz and in 2007/08 with 109 326 oz of gold at $367/oz. Production to June 2008 has been mainly from the M1 orebody, with the lowest stoping to date on the 580 m Level (615 m below surface) above development mining on the 520 Level. The rate of vertical advance of mining this orebody has averaged 150 vertical metres per year. The increased current combined JORC compliant resource and reserve position of over one million ounces, now includes the M2 and M3 orebodies with a current total average gold inventory up to 1400 - 1500 ovm. Exploration has intersected other shoots and all shoots are open at depth. The adjacent M2 and M3 orebodies are being developed for stoping. The narrow width of these zones from 1 m up to 4 m will require continued focus on ore zone definition, accurate development and careful drill and blast practice. In addition, to minimise cost increases in the current period of rapidly escalating labour, equipment, consumables and energy costs, high productivities will need to be maintained to allow the established rapid mining advance to continue at low unit costs.

Jan 1, 2008

Experimental Study of Ore Characterisation Methods for AG/SAG Mills Mining companies have long recognised the importance of ore breakage characterisation in design and optimisation of their crushing and grinding circuits. In one of the industry standard methods relevant to AG/SAG milling, ore is characterised using mainly high energy single impact tests. However recent DEM studies have suggested that breakage in AG/SAG mills is mainly due to low energy repetitive (or multiple) collisions rather than single high energy impacts. Hence a study was initiated at the JKMRC to investigate the behaviour of ores under such conditions. Carefully designed experiments have been performed using two laboratory mill sizes (0.6 m and 1.1 m diameter) to mimic the reported low energy repetitive impact breakage under low load conditions. The breakage characterisation parameters derived from standard high energy single impact tests on the same ores was then used to predict the low energy repetitive impact breakage. Significant bias was found in the simulations which clearly warranted further investigation. The extensive experimental program which followed employed three separate methods: the JKRBT (JK Rotary Breakage Tester) for high energy single impact and low energy repetitive impacts, gravity drops for low energy repetitive impacts, and laboratory mills to provide an extended number of low energy repetitive impacts. The breakage energies associated with these experiments were accurately measured with a high speed video camera. Two breakage indices, the degree of breakage (t10) and the breakage probability (S) were used to describe the breakage test results. This paper summarises the major results from this experimental study, which has facilitated the development of a new breakage characterization method relevant to AG/SAG mill modelling.

Sep 13, 2010

By C. J. Bauer

WHEN the pipe lines from the Middle East to the Mediterranean are completed, the Middle East supplies will relieve the strain on Western Hemisphere petroleum resources, part of which are now shipped for Eastern Hemisphere use. A sketch of the world petroleum situation as it appears today and in the immediate future will help in understanding the future place of Middle East oil supplies. Anticipated 1951 world demand is slightly under four billion barrels. To meet this requirement, crude output and refinery runs throughout the world (including Russia) must be increased about two million barrels per day above the average of 1947. This increment, together with pipe lines, tankers, and other facilities, will require in excess of ten billion dollars of new capital in the next four years. We are, therefore, looking forward to the greatest expansion program in the history of the industry.

Jan 1, 1948

By Roy C. Bartholomae

This Bureau of Mines report presents the results of a laboratory investigation of coal cutting mechanics and noise. These experiments were performed using linear cutting apparatus that operates over a broad cutting speed range. The influence of several coal cutting parameters on the noise, force, productivity, and specific energy associated with linear cuts was ascertained. Some basic theoretical aspects of coal cutting mechanics and noise generation are discussed, and the results of the laboratory experiments are used to formulate analytical models of the coal cutting forces and noise. The analytical model for coal cutting noise is then generalized to account for the more important effects of rotary cutting. Based on the generalized model, an estimate of the sound pressure level at an operator's position is made for a typical continuous mining machine. It is concluded that by employing linear rather than rotary cuts, using deeper depths of cut, slower cutting speeds, and more efficient cutting tools, it is possible to reduce the level of coal cutting noise, as well as provide benefits to other important areas of health and safety in underground mining.

Jan 1, 1983

By Paul W. Mayne

Axial pile performance can be rationally evaluated within an elastic continuum framework using field results from seismic piezocone tests (SCPTu). Adopting a versatile Randolph-type elastic pile model, the approach can be applied to either traditional top down loading using an anchored reaction beam or the newer Osterberg cell that simultaneously pushes the base and shaft in opposite directions. The axial load distribution within the shaft is also evaluated. For site-specific data at a given site, the SCPTu is an optimal means for collection of subsurface information because it combines penetrometer readings and downhole geophysics in one sounding. The results obtained are at opposite ends of the stress-strain-strength curves, specifically the peak strength for capacity interpretations and the small-strain stiffness (Emax) for evaluating the initial deformations. Axial pile capacity can be analyzed using both direct and indirect CPT methods. As such, the approach allows for a complete representation of the axial load-displacement-capacity curve for pile response. Case studies are presented for deep foundations situated in stiff clays at two national geotechnical test sites located in Houston and College Station, Texas, using top down loading, as well as a third case study of a drilled shaft in clay till loaded by O-cell in Alberta.

Jan 1, 2009

By G Corbett

Epithermal Au-Ag deposits form in the upper 1 km of the crust and are therefore best preserved in the less eroded Pacific rim arcs, although also recognised elsewhere. Two profoundly different ore fluid types give rise to low and high-sulfidation styles apparent as variations in ore and gangue mineralogy as well as characteristic zoned wall rock alteration. Low sulfidation epithermal Au deposits are derived from near neutral composite magmatic-meteoric fluids and divided into a group with a stronger intrusion association within arcs grading in time and space from: quartz-sulfide Au + Cu, to carbonate-base metal Au, and the epithermal quartz Au + Ag, whereas in strongly extensional settings the influx of meteoric waters promotes the development of chalcedony-ginguro banded Au-Ag epithermal veins. In the presence of continental crust particularly in Latin America polymetallic Ag-Au veins represent a Ag-rich equivalent of carbonate-base metal Au and passes upwards to chalcedony-ginguro Au-Ag style. By contrast high-sulfidation epithermal deposits are deposited from an acidic magmatic fluid which provided characteristic hydrothermal alteration by reaction with wall rocks grading outwards as mineral assemblages dominated by: vughy silica, alunite, pyrophyllite-diaspore, dickite-kaolin and then illite-smectite neutral clays. Au mineralisation is commonly refractory with enargite deposited dominantly later within the silica zone.Exploration for epithermal Au deposits benefits from an understanding of the controls to mineralisation which commonly give rise to the development of ore shoots with widest and highest Au grade ore zones. In low-sulfidation systems a combination of competent host rocks, dilatant structures, higher grade styles of mineralisation and efficient mechanism of Au deposition, contribute towards ore shoot formation. High-sulfidation epithermal Au systems occur in permeable volcanic and breccia host rocks which are linked by major structures to the magmatic source.CITATION:Corbett, G, 2013. Pacific Rim epithermal Au-Ag, in Proceedings World Gold 2013 , pp 5-14 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 26, 2013

By M. C. Harris

The past twenty to thirty years have seen unprecedented research activity aimed at understanding the performance of flotation froths. In 1990, Finch and Dobby coined the term froth recovery factor, Rr, to represent the efficiency of the froth in delivering particles from the pulp-froth interface to the concentrate. Originally defined for flotation columns, Rr has also been found useful in modeling mechanical flotation cells, and a number of techniques, based on quite different approaches, have been developed for measuring the froth recovery in industrial flotation cells. None of these techniques is universally successful. This paper reviews the concept of the froth recovery factor, and the basis of each of the techniques that have been proposed to measure Rr, and examines the reasons for this parameter being so difficult to pin down. Advances in Mineral Processing Science and Technology Proceedings of the 48th Annual Conference of Metallurgists of CIM Sudbury, Ontario, Canada Edited by

Jan 1, 2009

By Norman R. Paterson

"Under certain conditions useful gravity information of a reconnaissance nature can be obtained with-out measuring elevations by spirit level surveys on the ground. Gravity surveys can become part of a practical exploration program where high cost and slow progress previously hindered their utility. The advantages of the new elevation procedures are realized mainly in remote areas, in areas where ground transportation is slow and difficult and on surveys where the station spacing is large in comparison with the distance of the average optimum level shot. To utilize the advantages, transportation by helicopter or light plane is usually necessary. Methods of obtaining elevation control without leveling include: barometric altimeter survey; photogram-metric -APR spot determinations; methods involving the measurement of gravity gradients. The accuracies of these methods have been tested and compared with the normal accuracy of level surveys. It is felt that each method has its place in reconnaissance gravity work, particularly in the open areas of Northern Canada."

Jan 1, 1960

By J. J. Gilman

IT is well established that dislocations exist in crystals and account for crystal plasticity.' However, the origins of the dislocations are not clear. Among the means by which dislocations might originate are: a) through crystal growth accidents, b) by the collapse of condensed sheets of vacancies, or c) by stress-induced nucleation. Theoretical calculations' have indicated that the last means should require very high stresses in order to operate. The purpose of this paper is to show that the stresses around the tips of cracks are large enough to nucleate large numbers of dislocation loops. Specifically, it will be shown that (100) cleavage cracks in LiF crystals nucleate dislocations in regions of the crystals where dislocations did not exist previously. It is believed that the use of LiF for this investigation does not restrict the significance of the 1 results. In fact, it is believed that this complements and extends previous papers by the same author which dealt with metal crystals.10,14 The dislocations were detected by means of the etching techniques that were described in a previous paper by Gilman and Johnston.3 In that paper it

Jan 1, 1958