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By R. Fedors, T. Ghezzehei, D. Or

The phenomena of dripping into cavities (e.g. tunnels and drifts) within an unsaturated fractured porous medium under ventilated conditions is studied to improve estimates of dripping rates, drop sizes, and chemical composition of droplets that could affect the integrity of high-level nuclear waste packages placed in drifts. The sensitivity of dripping rates and water chemistry to ventilation is analyzed at various flow rates using analytical expressions linking flow on fracture surfaces to drop formation and evaporation. The model was tested using laboratory experiments involving machined aluminum slabs designed to represent a simplified morphology of a fracture face intersecting a cavity and using natural fracture surfaces from a welded tuff. Comparison with ten years of data on dripping in natural caves showed good agreement with model predictions. In particular, for low dripping rates, evaporation due to ventilation leads to larger drop sizes and alters drop chemistry by increasing the concentration of solutes in the drop relative to that in the water flowing on the fracture face.

Jan 1, 2000

By P Hagan, S Campbell

A study was undertaken using various methodologies to determine the optimum loading and hauling configuration for different haul routes at Callide Mine, a coal mine located in central Queensland. The study involved development of an effective flat haul model that was used to estimate production rates in truck and shovel circuits that employ two excavators of differing capacities then Queuing Theory to construct a capacity constraint model for scheduling purposes to assist in truck allocation and finally a cost model to determine the most cost effective fleet configuration.The model was calibrated using haul road design and performance data of trucks and excavators currently in use at the mine. The study examined the effects of cycle time for different truck and excavator fleet configurations on production rate and costs.The study found haulage cost was directly related to cycle time while production rate reduced with cycle time. The study indicates the approach taken in combining a cost model with a capacity constraint model offers a simple and effective method of scheduling an optimum number of trucks in a haulage circuit.CITATION:Campbell, S and Hagan, P, 2012. Selection of an optimum truck and shovel fleet based on effective flat haul and capacity constraint modelling, in Proceedings Eighth Open Pit OperatorsÆ Conference , pp 19-26 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 18, 2012

By Kim Le Roux

Cemented paste backfill (CPB) is a relatively new material that is increasingly being used in open stope mining across the world. The advantages offered by this high density tailings backfill include: a lower total cost associated with backfill placement, reduction of excess water underground; and the elimination of the need to process the tailings prior to placement. However, almost all backfill design is based on the performance of laboratory prepared samples since it is difficult to obtain undisturbed samples and conduct in situ testing due to the lack of access to backfilled stopes and the safety concerns in accessible areas. The logistical problems associated with conducting in situ testing and obtaining undisturbed samples of cemented paste backfill (CPB) has lead to the reliance in the mining industry on performance data obtained from laboratory prepared CPB samples. The degree to which these laboratory prepared samples represent the actual in situ condition is not well defined, partly because the influence of the mixing, transport and placement process on the material properties it is not well understood. Therefore, there is a need in the industry to understand how the material properties and hence the performance of the laboratory prepared CPB compares with in situ material. This paper reports the findings of an investigation into the in situ properties of CPB at a gold mine in Northern Ontario. A combination of in situ and laboratory testing, conducted on undisturbed samples, was used to quantify the in situ properties and performance of the CPB. These results were then compared to those of a laboratory based investigation of the same CPB. It is evident from this investigation that the bulk properties of the in situ CPB are different from those of the laboratory prepared CPB. The most significant observed differences are that the in situ CPB has a higher void ratio and a lower degree of saturation than the laboratory prepared material. The implication of these differences on the stability and performance of the CPB will also be explored in this paper. 1

May 1, 2002

By Thierry Bernard, Phillipe Cappello, Jean Marc Laboz

"The main reasons of the emergence of electronic detonators are linked to:-Local urban development close to mining operation sites-Environment regulations"

Jan 1, 1995

By Minggao Qian, Jialin Xu, Wenqi Lai, Weibin Zhu

This article advances an innovative concept explaining the connotation and technical system of "green" mining. The important theoretical foundations of "green" mining are the distribution and development characteristics of joints, fractures and bed separations, and the seepage and flow behavior of methane and water in the broken rock strata following coal mining. The main focus of "green" mining technique includes water-preservation in mining areas, coal mining under buildings with bed separation spaces grouted to retard surface subsidence, partial extraction and backfill mining, simultaneous extraction of coal and coal-bed methane, coal roadway supporting and underground disposal of rock waste, underground coal gasification, etc.

Jan 1, 2005

By Michael G. Zabetakis

Vaporization of a solvent into an airstream at temperatures above the flash point of the solvent is hazardous if vapor is liberated at a rate that is sufficient to produce a flammable mixture. In may operations, the formation of such a mixture is prevented by rapid diluton of the vapors with air. The minimum air requirements for the dilution process are determined by the composition of the lower limit mixture at the temperature and pressure of interest. Because limits of flammability cannot always be calculated with sufficient accuracy for a proposed operation, they .rust often be determined experimentally. The present Bureau of Mines investigation was prompted by an application in which the lower limits of flammability of a number of solvent mixtures were required over a range of temperatures at atmospheric pressure. These limits were to be used along with available solvent-vaporization-rate data to establish minimum airflow rates needed to prevent the formation of flammable mixtures in an industrial process. Lower limits of flammability were determined for a number of mixtures containing methyl ethyl ketone, tetrahydrofuran, toluene, ethyl alcohol, and ethyl acetate in dry air-at atmospheric pressure and temperatures between 16° and 200° C. The experimental values were found to be in good agreement at with those predicted by Le Chatelier's law; variations with temperature were in fair agreement with those predicted by a modified Burgess-Wheeler law.

Jan 1, 1962

By Enrique Hermana, Raymond F. Dalton, Raymond Price, Barry Hoffman

Introduction Chloride leaching of sulfidic ores has been known about for 100 years. It is recognized as a highly efficient process to produce concentrated leach solutions. Despite this, chloride leaching found little practical application until recently because of the highly corrosive nature of concentrated chloride solutions. This situation changed with the advent of modern construction materials such as GRP, polypropylene, butyl rubber, etc. And in the past 20 years, research has burgeoned in the area of chloride hydrometallurgy. The research impetus stemmed from the belief that chloride leaching can provide an alternative to smelters for treating conventional copper concentrates. In addition, chloride leaching was thought to be suitable for on-site refining of copper at small tonnage operations located in remote areas. And chloride leaching was seen as a logical choice for treating unusual or complex concentrates that smelters could not handle. Underpinning these beliefs is the attraction of chloride-based hydrometallurgical processes for sulfidic ores. Sulfur is liberated in the environmentally acceptable elemental form rather than as sulfur dioxide. Consequently, sulfidic copper ores have received much attention. A review of Edmiston in 1984 describes 12 chloride-based processes. All managed to progress past the "laboratory beaker" stage of evaluation. In nine of these processes, copper is recovered electrolytically. Key features of these processes are summarized in Table 1. Table 1 shows that the processes must include a copper solvent-extraction step. Only these processes produce a clean copper electrolyte capable, with no further refining, of directly producing high quality cathode copper. The ICI/Elkem process is the exception here. The rest of these processes use o-hydroxyaryloxime reagents for the solvent extraction of copper and are constrained by the performance of these reagents. This is governed by hydrogen ion concentration (Equation 1). The practical result is that only relatively dilute copper solutions can be treated (Canmet, Phelps Dodge), or careful pH control or adjustment is necessary (Minimet Recherche, B.H.A.S.). This negates one of the attractive features of chloride leaching - its ability to produce very concentrated copper solutions, up to 40-45 gl-1. [ ] In all four processes using o¬hydroxyaryloxime extractants, solvent extraction transfers the copper from the chloride to the sulfate medium. To minimize cross-contamination of the two anions, the loaded organic solvent has to be scrubbed to remove the chloride ion. This represents an additional complication in the solvent extraction circuit. And it

Jan 1, 1988

By M. Kolahdoozan, H. Noori

"Molybdenite flotation at the Sarcheshmeh Copper Complex has been studied as a function of pulp potential controlled using two types of sodium sulfide (mineral and chemical origin). Air and either plant or laboratory nitrogen was used as a flotation carrier. Potential measurements were taken in the rougher cells as it was proved to be vital for molybdenite flotation. Flotation test with reducing potentials between -525 and -575 mv gave the highest molybdenum recovery (93%) and the lowest copper recovery of between 4 to 15% can be achieved. These potentials did not change considerably by adding increasing amounts of sodium sulfide. At more oxidizing potentials, i.e. more than -450 mv, the recovery of copper increases while the recovery of Molybdenite drops.INTRODUCTIONThe average copper ore feed grade to the Sarcheshmeh flotation circuit is typically set at about 1.2% Cu. to be delivered to processing unit is set at 1.2%. There is no control of molybdenum (Mo) which usually varies between 0.01 and 0.04%. During concentration, a Cu-Mo concentrate is produced by one stage of roughing, one stage of scavenging and two stages of cleaning. Final concentrate contains 0.5% of Mo and 32% of Cu while the particle size is 90% under 42 microns (325 mesh). Because of the high grade of Mo in this concentrate and economic interests, a molybdenite flotation circuit plant was built in 1984. This plant was designed to treat 1432 tons of Cu-Mo concentrate (on dry basis) and consists of one stage of roughing followed by seven stages of cleaning. After dewatering the concentrate contains 53.7% Mo which is dried and shipped overseas.The majority of molybdenite production worldwide comes as byproduct of copper porphyry ores. The Mo grade typically varies between 0.01 and 0.06% and accompanies chalcopyrite and chalcocite. Due to the hydrophobic nature of molybdenite, it can be floated along with copper sulfides. Thus the overall concentrate of Cu and Mo contains 0.2-1% of molybdenum as well as 20-42% of copper. To separate the Mo, Cu and Fe minerals, depressants are used. The most important of which are sodium sulfide, sodium hydrosulfide, Anomal D, sodium cyanide, ferropottasium and ferro-cyanide. The first three regeants have been extensively used for successfully treating porphyry ores [1]. In the molybdenum plant of the Sarcheshmeh Copper Complex only sodium sulfide has been used to date."

Jan 1, 2006

By L. C. IlsLey

"This paper is intended for State mining inspectors, safety engineers of mining companies, and others interested in electrical safety inspection in mines, It presents the important points that should be watched by inspectors, and briefly reviews the work of the Bureau of Mines in the testing of electrical apparatus or equipment to determine their permissibility for use in gaseous coal mines.Important Points in Inspecting Mine Electrical Installations or Equipment.The following questions are intended as suggestions to a safety inspector who may not have been specially trained in looking for defects in installations or in the maintenance of electrical equipment:1.,Is there someone in charge of electrical equipment who is fitted for this position by ability, training and experience?2.,Is a systematic recorded inspection of electrical equipment made at least once each month?3.,Is a plan map of the mine kept showing all permanently installed electrical machinery and apparatus, including cables, conductors light etc.?4.,Are there instructions for resuscitation from electric shock posted in each surface and underground electrical station, also at the entrance of the mine and in the first aid station and are employees working on electrical equipment trained to apply these instructions in case of accident?"

Nov 1, 1923

By J. A. Finch, T. Weber

"Bubble size in flotation is strongly influenced by frother concentration in the pulp liquor. There is no accepted method of judging the dosage and thus how effectively frother is being distributed to control bubble size is not known. This will hamper certain plant test work where constant bubble size (e.g., when comparing reagents) is implicit. Bubble size is not easy to measure in plant and gas holdup is suggested as a surrogate in the development of a proposed “meter” to estimate frother concentration. The meter is calibrated to enable frother concentration to be estimated on samples drawn from any location in the flotation circuit. The meter and its calibration are described.As general background the action of frothers in reducing bubble size by retarding coalescence is discussed and the bubble generation systems for which the effect will be important are considered.INTRODUCTIONBubble size is an important variable in flotation. The flotation rate constant k has been related to mean bubble size Db by k % Db -n with n ranging up to 2 (Ahmed and Jameson, 1989) and even 3 (Yoon, 1993). Recent work has introduced k % Sb, the bubble surface area flux, given by Sb = 6 Jg / Db (where Jg is the superficial gas velocity or volumetric gas rate per unit cell cross-sectional area), i.e., n = 1 (Gorain et al., 1998). Regardless of the exact value of n, the significance of bubble size in flotation remains. In many plant trials it will be implicit that bubble size has not changed and attention will focus on other process variables. The measurement of bubble size, or a surrogate, would be useful addition to any test work.The size of the bubbles is influenced by several variables, gas rate, gas injection system (e.g., impeller design and rotation speed in a mechanical cell) and pulp composition (chemicals added, percent solids), but the most important is frother concentration (Harris, 1976). The frother functions by retarding coalescence (see appendix), which acts to preserve the small bubbles generated in the pulp (and permits a stable froth to form on top of the pulp). The measurement of bubble size is thus related to measuring frother concentration. The distribution of frother and whether sufficient (to control bubble size) is reaching the intended parts of the circuit then become important considerations. At present there is no simple way to judge frother concentration in a plant. In this communication we introduce a possible methodology, a “frother concentration meter”."

Jan 1, 2003

This paper gives a general description of mining operations at the Mount Morgan Open Cut. Certain features such as blast hole diamond drilling, which may be of particular interest, have been described in detail. The paper is covered by the following headings:-1. History2. Ore Reserves and Production Data3. Mining Methodsa. Design and Layout of the Open Cut.b. Overburden and Ore Breaking.c. Loading and Transport of Ore and Overburden.d. Transport of Men and Supplies.e. Survey and Assay.f. Drainage.4. Power and Compressed Air5. Mining Equipmenta. Shovels.b. Mine Transport.c. Dozer Tractors.d. Drill Shop.6. CostsHISTORYIn order to appreciate fully the design of the present Open Cut at Mount Morgan, a brief outline of the Mine's history and past mining methods is necessary. In 1882 the Mount Morgan Syndicate began to mine this large gold-copper orebody by small hill-top quarries...

Jan 1, 1952

By R. J. Durrheim

The seismicity of the Witwatersrand Basin and Bushveld Complex has been analysed, revealing significant differences in source and damage mechanisms. A comparison of mining taking place at similar depths and extraction ratios indicates that the seismic energy release rate is considerably lower in the Bushveld Complex. Technologies introduced in the past decade to manage the risk of rockbursting are reviewed and assessed. These include geophysical methods to detect hazardous structures ahead of mining, new mining layouts and support systems, and the mechanization of in-stope processes. Despite important differences in geology, in situ stress, mining layouts and methods, there are many opportunities for the reciprocal transfer of knowledge, experience and technology between the gold and platinum mining sectors to improve the management of the rockburst risk.

Jan 1, 2005

By C. Birch

"A mineral resource royalty is payment to the holder of mineral rights for the utilization of the mineral resource. In South Africa, this payment is made to the State as holder of the mineral rights (The Mineral and Petroleum Resources Royalty Act of 2008). The principle purpose of this research paper is to identify if the State is benefitting from the mineral resource royalty by considering its impact on seven individual Witwatersrand gold mines.For this study, a simple financial optimiser model was created in Microsoft Excel that links the ore flow, block listing and the cash flow (excluding or including the cost of the mineral resource royalty). Mixed integer linear programing (the Excel Solver function) is utilised to optimise either profit or NPV (at 9% and 12%) by adjusting the cut-off grade.The impact on each of the seven mines mine was different but overall R7.9 billion is estimated to be paid in mineral resource royalty over their expected remaining lives. Due to the cost of the mineral resource royalty and increasing the cut-off grades, the total revenue decreases by R10 billion. A significant portion of this lost revenue would have been paid to the State in the forms of other taxation including company income tax which decreases by R2.8 billion. It is recommended that an industry wide investigation be conducted to determine if the resource royalty is adding to the State’s revenue, or destroying value including premature job losses. IntroductionTaxation of mining companies has become a very topical subject in the context of the postapartheid South African economy. There have been increasing calls for equitable re-distribution of mineral asset wealth for the benefit of all South African citizens and not just to an elite few. As part of the Mineral and Petroleum Resources Development Act (MRDPA) of 2002, the Mineral and Petroleum Resources Royalty Act of 2008 was introduced and came into effect in 2010 (South African Revenue Services, 2008). A mineral royalty is payment to the holder of mineral rights for the utilization of the mineral resource. In the case where the holder of the mineral right is the State, then this payment is made to the State (Cawood, 1999). Studies on the potential impact of the mineral resource royalty on the minerals industry have been conducted during the formulation of the Act and subesquently. These include Cawood and Macfarlane (2003), who suggested a cap of no more than 3% (Cawood and Macfarlane, 2003), as well as Cawood (2011), who predicted rising cut-off grades and thus reducing mineral reserves. In 2013, Grobler completed a Masters in the Faculty of Economic and Management Sciences at the University of Pretoria and concluded that the impact of the mineral resource royalty on the economy is positive. She stated that the royalty has caused operating expenses to increase, but this does not appear to be significant. Furthermore, she stated that the royalty has not affected the way the mining industry operates (Grobler, 2014)."

Jan 1, 2016

By David K. Ingram

The Bureau of Mines conducted underground mapping and rock strength tests to describe and analyze mine roof conditions surrounding ancient stream channel deposits (paleochannels) in the Pittsburgh Coalbed in southwestern Pennsylvania. Paleochannels in the study mine consist of sandstone and/or siltstone and affect the Pittsburgh Coalbed through erosion and/or differential compaction. Differentially compacted sediments within and adjacent to channel deposits caused slip planes, faults, clay-dike faults, clastic dikes (clay veins), coalbed rolls, and slumped structures. Paleochannels can be predicted by recognizing these features and associated sediments, which allows for modification of long-and short-term mine planning and development. Mine entries located beneath the paleochannel deposits exhibit less stable roof conditions than entries with normal (nonchannel) roof. Paleochannel deposits comprise only one-fourth of the mapped study area; however, they contain one-half of the hazardous roof. This study deter-mined that rock strength evaluations, including rock quality designation (RQD), unconfined uniaxial compressive strength testing, and point-load testing, must be accompanied by underground observations. Suggested remedial support techniques include angled and/or longer tensioned bolts, steel mats or crossbars, steel sets or cribbing, and roof trusses.

Jan 1, 1987

By Ike Isagon

Backfilling has become an integral process in underground mining operations, with varying cost-effective forms of application. This study is a part of an overall investigation into the backfill practices at Copper Cliff North Mine (CCNM). The mine is well-diversified in terms of backfill methodologies and practices. The mine employs different fill methods such as cemented and un-cemented rock fill, slag fill, and cemented hydraulic fill (or a combination of them) in order to fill the voids created by mining activities. The history of backfilling at CCNM has been dominated by cemented and un-cemented slag fill. Each fill method requires operating and engineering guidelines for a cost effective quality control system. The purpose of this paper is to briefly discuss the backfill methods and applications, operating guidelines, and engineering requirements at CCNM.

May 1, 2011

By Ron Board

"Probably the only thing that can he said about conveyor belt scales without fear of contradiction is that they have been responsible for as much or more puzzlement, argument and frustration than almost any other piece of equipment commonly found in ore processing facilities. To be sure, there are certainly a good number of satisfactory installations, some of these having been achieved through forethought and careful consideration in the early stages of plant design, while others seem to owl their success to what seems to be little more than good luck or white witchcraft; and it might also appear that black witchcraft or other gremlins are solely responsible for the poor performance of many belt scales that enjoy well-designed installations with apparently good operating conditions.This paper is intended to acquaint users with the major operational reasons why belt scales seem so fickle (and downright ornery too), and establish some reasonably simple test standards against which the performance of scale hardware can be measured. Then it shall be shown how the design flexibility of electronic scale systems can offer particular advantages in overcoming some of these everyday problems in a practical way, and without undue inconvenience to plant operations."

Jan 1, 1975

By Staff

The mineral sector is a fundamental part of the U.S. economy. It contributes to the real gross domestic product (GDP) at several stages—basic (mining), intermediate processing, manufacturing of finished goods and construction. The estimated growth rate for the real GDP of the United States for 2006 was 3.2 percent. The nominal GDP was about $13.2 trillion. The value of minerals mined in the United States in 2006 rose significantly because of increased unit price for some metals—particularly copper, zinc and other base metals. There was also a significant reduction in metals stocks. Infrastructure expansion and manufacturing in China and India have been absorbing a significant portion of world output. Domestic production increased for some industrial mineral materials, mainly cement and construction sand and gravel. Production of crushed stone declined less than 1 percent in 2006 compared

Jan 1, 2007

By JAMES BRADLEY

THE Spanish mine is in Nevada County, California, 21 miles northeast of Nevada City by road, and 3 miles north of the town of Washington. The mill and surface buildings are on Poorman's Creek at an ele¬vation of 3,000 ft. About 35 men are employed at the property, which is producing 50 tons of ore per day. Electric power, purchased from the Pacific Gas and Electric Co., is brought in from Alleghany, seven miles away. In the early days the mine was operated by F. W. and P. R. Bradley. Between 1883 and 1899 the production was approximately $900,0)0 in gold. The ore treated during; that time yielded an average of 60 cents per ton and was mined and milled at a profit. The surface, oxidized, ore deposit was 30 to 100 ft. wide and was mined by the glory-hole method. The milling-plant at first consisted of a Blake crusher an-1 a small Huntington mill; later three large Huntington mills were added to the plant. In 1893 a new mill was built at a lower adit- It consisted of 10 stamps and four Huntington mills. The gold was secured by amalgamation, the recovery being about 50 per rent.

Jan 1, 1931

By Cameron J. Clayton

Cumberland Resources Ltd. is currently planning to develop the Meadowbank Gold Project located some 70 km north of Baker Lake, Nunavut in the eastern Canadian Arctic region. The Meadowbank Gold Project currently consists of up to 6 main gold bearing deposits within relatively close proximity to one another which are planned to be mined primarily as a truck-and-shovel open pit operation and using conventional gold milling and extraction. The mining plan indicates that approximately 20 million tonnes of ore will be mined over a mine life of 8 to 10 years. The operation will generate approximately 160 million tonnes of mine waste rock and about 20 million tonnes of tailings. The site is located in an area of continuous permafrost with mean annual average air temperature of -11 °C and average annual precipitation in the range of 100 to 200 mm. Lakes cover a large portion of the site, including some of the ore deposits. The location of the site, and the physical environment, pose challenges to development. This paper presents some of the designs considered to effectively manage and utilize the waste products of mining at the project.

May 1, 2004

By M. Bryson, S. Broom-Fendley, A. E. Brady, W. Dawes, G. Beer, N. Sehlotho, S. Al-Ali, F. Wall

The Songwe Hill Rare Earth Element (REE) Project in Malawi is a near surface carbonatite-hosted deposit with synchysite and apatite as the key REE bearing minerals. A strong focus has been placed on understanding the REE mineralogy of the Songwe Hill carbonatite as REE mineralogical characterization is a crucial component at all steps, ranging from exploration to production, in the development of REE deposits. Extensive mineralogical work has been completed at Camborne School of Mines (CSM), Natural History Museum (NHM) London, British Geological Survey (BGS), Nottingham, Aberystwyth University, Wales, Mintek, South Africa, and SGS Inc. Canada (SGS). The following analytical techniques were used: scanning electron microscopy (SEM), electron microprobe analysis (EMPA), laser ablation inductively coupled plasma mass spectrometry (LA-ICP-MS), and QEMSCAN®. The detailed studies have demonstrated that while synchysite is light-REE enriched as expected, apatite, abundant in all stages of carbonatite emplacement at Songwe, is anomalously enriched in mid- to heavy-REE relative to apatite in other carbonatite deposits. The distribution of REE within synchysite and apatite at Songwe has had important implications for the direction of the metallurgical work with respect to the behaviour of light and heavy REE through the flotation and hydrometallurgical stages of the flow sheet.

Jan 1, 2014