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The final criterion of the payability of gold production must necessarily be the cost per ounce. While the cost per ton or per cubic yard is a convenient comparative measure, and a necessary one for operating control, it is obviously insufficient as a means of judging whether or not, in a given set of circumstances, a gold production operation can be carried out economically.. An alternative method, which brings in the ton or cubic yard as a unit of measurement, is to determine what the cost per ton or per cubic yard is equivalent to, in terms of pennyweights or grains at the ruling price of gold. But this is really only another way of stating that the cost per unit of product is the ultimate measure of profit or loss, and not the cost per unit of input material.It is accordingly of interest to determine the cost per ounce of Australian gold production and to compare it with the cost of gold production in some other parts of the world which are important gold producers. The attached Appendix, which it, is believed comprises practically all Australian mines of any consequence (except Big Bell and the de Bernales group, for which. insufficient information is available), shows this information for financial years ending between 31st December, 1938, and 31st December, 1939, the production of gold dealt with being some 1,084,000 oz., or 66% of the total Australian production for 1939. For comparative purposes the cost of producing gold at representative mines in New Guinea, New Zealand, Fiji, Canada, India, West Africa and Rhodesia, and the average cost of production at the Rand mines, has also been included.It is obviously necessary in a study of this sort that the basis of the cost should be clearly set out and should be as nearly as possible uniform. The "cost" as calculated herein comprises the following items:

Jan 1, 1940

THE Australian States and New Zealand have for many years provided instruction in mining, metallurgy, geology, and their allied subjects through their Universities and Schools of Mines. Though such education has been provided at considerable expenditure, there is no legal recognition whatever given to the holders of degrees, diplomas, or certificates granted by the authorities controlling these educational institutions. Certain of the, Australian States and New Zealand require, however, that mine managers and other officials employed by the mining industry shall 'hold' certificates granted by their respective Mines Departments.The position at present is that each educational institution is practically a law unto itself, and therefore in many ways there is a want of uniformity amongst institutions all aiming at the same result the training of mining men. The object of this paper is, in the first place, to point out certain aspects of this want of uniformity, and thereby how disabilities accrue. Some points in connection with the granting of State certificates will also be...

Jan 1, 1915

By Robert L. Lagace, Michael A. Grossman, Alfred G. Emslie

This supplement to the final report is a collection of interim and monthly reports and working memoranda prepared during the course of this program to document the progress, methods and results of the work. The supplement is divided into the self-contained sections briefly described below. I. SELECTED MONTHLY TECHNICAL LETTER REPORTS - This section is a collection of selected monthly technical letter reports that briefly summarizes the chronological development of the analyses and results, including the different concepts, theoretical approaches, and methods of data reduction and analysis explored, and the discoveries, insights and findings made along the way. More detailed treatments of the key developments and results are presented in the interim reports and working memoranda reproduced in subsequent sections of this supplementary report, and in the separately published final report. II. BACKGROUND THEORY FOR MEASUREMENT PROGRAM ON MEDIUM AND HIGH FREQUENCY RADIO TRANSMISSION IN COAL SEAMS - Working Memorandum - This working memorandum gives the initial background theory needed to guide the planned measurements and data analyses of this program to describe the propagation of radio waves-in coal seams for the frequency range of 50 to 5000 kHz. III. MODELLING AND DATA ANALYSIS OF IN-MINE ELECTROMAGNETIC WAVE PROPAGATION - Interim Report - This interim report presents detailed treatments of the data analysis and modelling techniques used, and the associated results obtained from conductor-free area propagation data measured by T. Cory in the first set of six coal mines. The mines were located in three highcoal seams; the Pittsburgh seam in northern West Virginia, the Pocahontas No. 3 seam in Virginia, and the Herrin No. 6 seam in Illinois. IV. ANALYSIS OF MF PROPAGATION DATA FROM MARGARET NO. 11, NANTY GLO, EHRENFELD, AND ADRIAN COAL MINES - Interim Report - This interim report presents detailed treatments of several improved data analysis and Modelling methods applied to, and results obtained from, conductor-free area propagation data measured by T. Cory in the second set of four coal mines. These mines were located in three low-coal seams; the Upper Freeport seam in Pennsylvania and West Virginia and the Lower Freeport and Lower Kittaning seams in Pennsylvania. In addition, this report contains a reprint of two papers presented at a Bureau of Mines sponsored EM Guided Waves in Mine Environments Workshop held in Boulder Colorado. One paper describes the three-layer theoretical model and its application to the conductor-free data from the first six mines, while the second paper describes the theory developed for the coupling of loop antennas to a cable in a mine tunnel in a coal seam, and the model's application to data in the vicinity of cables taken at the Margaret No. 11 mine. V. A METHOD FOR NONINTRUSIVE, IN-SITU MEASUREMENT OF COAL AND ROCK CONDUCTIVITIES IN A COAL MINE TUNNEL - Working Memorandum - This working memorandum briefly describes a possible direct method, and its potential advantages and shortcomings, for performing in-situ measurements of coal and rock conductivities. VI. MODELLING AND DATA ANALYSIS OF IN-MINE ELECTROMAGNETIC WAVE PROPAGATION FROM 50 to 5000 kHz - Interim Report - This interim report presents a summary comparison of theoretical and experimental results for conductor- free areas in eleven coal mines located in seven coal seams. It also summarizes the findings and their implications for intrinsically-safe, portable radio communications between roving miners.

Jan 2, 1980

By Barnard Foo

Topics of Discussion ?Location ?Production & Mining Operation ?Mineralization ?Mining Method ?Rock Mechanics Issues ?Ground Control Problems ?Numerical Analysis ?RockmassBehaviour ?Conclusions & Recommendations ?Discussion

Feb 1, 2008

By C. Bise, A. Seal

Excessive noise has long been a hazard in the coal-mining industry. Studies conducted during both the mid-1970s and the mid-1990s consistently show that hearing hearing loss within the mining industry persists, in spite of regulatory requirements and sampling technology advances. When MSHA’s new health standards to protect miners from hearing loss took effect in September 2000, the potential developed to re-evaluate sampling approaches with regard to noise-source hazards, such as task-based methods. This paper describes results from a Penn State research project currently being conducted at an underground coal mine. Each occupation was reviewed, and a list of tasks or processes performed by each was generated. Traditional personal noise dosimetry, static sound pressure levels, and equipment sound mappings have been performed. Job tasks indicating the greatest levels of daily noise exposure and contributions to noise dose have been identified and will be discussed.

Jan 1, 2001

By David Newman

Multiple seam underground coal mining is ubiquitous in Appalachia. Many coal properties have from two to ten or more economically mineable coal seams. Because of the ability to penetrate 600-feet to 1,000-feet or deeper into the mountainside leaving panels of web pillars separated by barrier pillars, highwall mining is a hybrid between surface and underground mining. As a result, many of the ground control problems commonly associated with multiple seam underground mining exist with multiple seam highwall mining. These include: i) superjacent mines, subsidence and caving propagating to overlying seams and ii) mines, pooled water and the potential for inundation resulting from a iii) breakthrough between an active highwall mine and abandoned underground mines in the same or superjacent seams, the presence of overlying or underlying mine workings where iv) pillar columnization in the active mine is not possible, and v) the question of whether to mine ?top-down? or ?bottom-up.? Case histories are presented from southeast Ohio where the Middle Kittanning No. 6 and Lower Kittanning No. 5 seams are successfully highwall mined with 20-feet to 40-feet of largely shale interburden. A highwall miner was caught in a highwall collapse in the Middle Kittanning No. 6 seam when it was overmining auger workings in the Lower Kittanning No. 5 seam. Multiple seam interaction caused by the overmining transferred overburden stress onto the Lower Kittanning No. 5 seam. The overburden concentrated through web pillars initiated auger web pillar failure, resulting in roof collapse and subsidence in the Lower Kittanning No. 5 seam that propagated up to the Middle Kittanning No. 6 seam. A second case history involves multiple seam highwall mining in southeast Kentucky. The A3 Rider, A3, B2, and C1 seams are being highwall mined around abandoned underground mines in the A3 and B2 seams and augering in the A3 Rider. Because of spoil storage constraints for the surface contour mines, the highwall miner follows the active pits in each seam so a standard sequence of ?top-down? or ?bottom-up mining does not exist. The planning and engineering design considerations associated with each case history is presented along with geotechnical information. In the case of the highwall collapse a back analysis is presented of the stability and safety factors for the auger webs, highwall miner web pillars, and barrier pillars.

Jan 1, 2009

By D. J. Duck

INTRODUCTION Construction of the 3,600,000 kilowatt Third Powerplant at Grand Coulee Dam started in late summer 1967 approximately one year after authorization by Congress. The project was made possible by a treaty with Canada providing for cooperative development of the Columbia River which was signed in 1961 and ratified in 1964. Description of the Work The existing Grand Coulee power complex consisted of a dam and two powerplants with a switchyard serving each plant, one on each bank of the river. See Fig. 1. The switching facilities were connected to transformers at the powerplants by overhead 230-kilovolt and 115-kilovolt tie circuit conductors. [ ]

Jan 1, 1997

By Roy Peratrovich

An innovative cost-saving design for a 1,100 ft bulkhead on the Tanana River at Nenana in central Alaska won Alaska Construction & Oil magazine's Construction Achievement Award. The designers utilized materials available nearby at low cost. They planned sheet pile enclosures, self-anchored back into shore without cell closures. They made it practical for inexperienced local workers to do the building. And they completed the project on time and below a reduced budget.

Jan 1, 1985

By G. L. Dollinger, J. P. Kindsvater, J. A. Gibson

DESIGN AND DEVELOPMENT OF A MOBILE MINING MACHINE INTRODUCTION Until recently, developing tunnels in hard rock has been accomplished either by the drill and blast method or by using large tunnel boring machines. Combining the flexibility of drill and blast with the safety, efficiency and productivity of a TBM has been the objective of mining machinery producers for many years. In 1979 The Robbins Company began developing a new type of tunnel boring machine called the Mobile Miner to meet these objectives. The development and testing of the Mobile Miner is the subject of this paper. CURRENT MINE DEVELOPMENT METHODS Both the drill and blast method and the use of a TBM have notable advantages and disadvantages. The drill and blast method, for example, allows much more flexibility, requires less setup time and has a lower initial cost than a TBM; however it requires the use of high explosives and blasting has a disrupting effect on the rock surrounding the development. A TBM, on the other hand, is safer to operate, more efficient, more productive and produces smooth tunnel walls without disturbing the surrounding rock. In longer straight drifts it will easily out distance the drill and blast method. The TBM, however, is limited to boring a circular cross section tunnel which must then be modified if a flat floor is required. It also has a higher initial cost and an extended setup period.

Jan 1, 1985

By Philip Bonanno

The first major project in Boston's new access to Logan Airport was an Immersed Tube Tunnel which extended 3,600 ft. from South Boston to East Boston across Boston Harbor. Twelve tube sections, 300 ft. long x 80 ft. wide x 40 ft. high, were lowered into an excavated trench which was then backfilled. The depth of the trench at mid-channel was approximately 100 ft. below mean high tide. At the East Boston end the tube terminated close to the present shoreline. At the South Boston end the tube terminated approximately 600 ft. inland from the shore line where a ventilation building was required. The depth to subgrade from existing ground at the ventilation structure was approximately 85 ft. The soil profile at this location consisted of 50 ft. of miscellaneous rubble fill, 25 ft. of Boston blue clay, and 10 ft. of glacial till. A small amount of weathered argellite existed at the north end of the structure. Part of the Morrison Knudsen, Interbeton, J.F. White, Joint Venture contract to construct the Immersed Tube Tunnel (ITT) was the construction of a cofferdam at the tube/ventilation building interface. The ventilation building was to be built as part of another contract. The construction of the cofferdam and the excavation to the building subgrade was the responsibility of J.F. White Contracting Company.

Jan 1, 1994

By Douglas D. Campbell, Alan S. Imrie

INTRODUCTION Mica Dam is located on the Columbia River just downstream of the "big bend" of the Columbia, where it leaves the Rocky Mountain Trench and begins to flow southwards, approximately 85 miles north of Revelstoke, B.C. (Fig.l). The dam was built as one of three Columbia River Treaty dams in British Columbia which regulate the flow of the river and provide flood control downstream in Washington and Oregon. In addition, the three Canadian dams make possible greatly increased power generation at existing power plants in the United States and further development along the Columbia and its tributaries in British Columbia. History Mica Dam, constructed during the period 1967 to 1973, is a zoned earthfill embankment 650 feet high, which has 12,000,000 acre feet of live storage and an available head of 550 feet. Power intakes which would serve either a six-unit surface or underground plant were incorporated into the design and construction of the right abutment of the dam (Fig. 2). Prior to 1971 alternative 2500 MW capacity surface and underground powerplant arrangements were studied. Cost estimates for the two alternatives were the same within the limits of estimating procedures. However, it was believed there was a certain continuing risk in locating a surface plant at the toe of a steep canyon wall. Therefore, in 1971 a tentative decision was made to proceed with an underground plant, provided that additional exploration confirmed assumptions of

Jan 1, 1976

Experimental investigations were made into the decoppering of lead, tin, and bismuth using a technique which approximates on the small scale to industrial practice.Decoppering lead with sulphur alone in the absence of tin and silver did not proceed to low copper values. With a minimum of 0·02 per cent Ag or 0·10 per cent Sn, lead was decoppered to less than 0·01 per cent Cuby stirring with 0·2 per cent S over 10 min. Decoppering of tin or bismuth proceeded rapidly to very low levels in the absence of any "catalytic" impurities, at any rate in excess of those concentrations normally encountered in commercially pure tin or bismuth.

Jan 1, 1971

Ore from the Hill 50 Gold Mine, consists essentially of pyrrhotite, with minor amounts of pyrite, chalcopyrite, galena and gold, replacing "jaspilite." The gold is associated with the sulphides as shotty particles. About 8 per cent. of the particles are less than '0.010'mm.The Hill 50 Gold Mine, at Boogardie in the Murchison goldfield, about 300 miles north-west of Kalgoorlie, was first operated in 1898. From 1898 to 1917 it yielded 20,000 short tons of ore for a recovery of 6,225' fine oz. gold, equivalent to a recovery grade of 6.22 dwt. gold per ton.The present company since the beginning of its operations in September 1936 to June 30, 1954, produced 811,081 short tons of ore for a recovery of 242,554 fine oz. of gold, equivalent to a recovery grade of 5.98 dwt. gold per ton.The grade of ore (recovery grade) has shown a remarkable variation with depth. From the 120 ft. Ievel to the 300 ft. level, it averaged about 7.5 dwt. Au per ton. Below the 300 ft. level it declined, and from 300 ft. to 400 ft. averaged 4.9 dwt. per ton, falling further with increasing depth to about 4.5 dwt. per ton at the '600 ft. level.On the 820 ft. level the grade improved sharply to about 14 dwt: per ton, and this grade has persisted down to and on the 1060 ft. level.The orebody has been described by Finucane (1953) as a series of lenses of finely banded jaspilite conformably enclosed by greenstones. The jaspilite is tightly folded and faulted and intruded by dykes of felspar porpnyry.Mineralization is practically restricted to the jaspilite.

Jan 1, 1955

By William Paxton Hewitt

In the Santa Eulalia mining district, Chihuahua, Mexico, the main mineral zone occurs in a gently warped dome with simple associated structures and contains orebodies distributed through a vertical range of over 900 m. The stratigraphy of 1945 m of geologic column is described and an abrupt shrinkage is documented in the volume of limestones immediately associated with large orebodies. Termed collapse, this shrinkage occurred during the course of mineralization. Contacts of limestone with igneous sills, and 366 m of fossiliferous limestone beds, particularly the main fossil bed, have been important localizers of ore-bodies. Up-dip gravitational controls may have been important. Fractures have exerted only local controls and over short distances. Thrust zones have displaced diabase sills and in turn have been intruded by upper members of a group of rhyolite porphyry sills. Mineralizers were related to the rhyolite porphyry, and ore has been introduced to the limestone where certain sills lie between the displaced segments of diabase and have been cut by north-south fissures. Speculations are raised that fields of unreleased strain may have influenced orebody locations; the district's zoning pattern may be a reflection of the filtering effects of large volumes of limestone; and that metal-rich brines mobilized by or incorporated within the invading porphyry might be the source of the mineralizers. PRODUCTION AND LOCATION The Santa Eulalia district has a production record approaching or exceeding 35 million tons of silver-lead-zinc ore from replacement bodies in limestone. In operation since the first years of the 17001s, production has been confined to two intensely mineralized areas separated by 2 km of barren limestone, the West Camp and the East Camp. Probably 50% has been mined since 1908, with over 90% from the West Camp. The West Camp, exploited through a vertical depth of 1945 m and now approaching exhaustion, is the subject of this paper. The district lies in the northern part of the Santa Eulalia Mountains in central Chihuahua in northern Mexico, 400 km south of El Paso, Texas, and 22 km southeast of Chihuahua City. Dominating the northern part of the range is the Cerro de Santa Eulalia, a broad rolling upland which contains the district's West and East Camps. The West Camp, on the west side of the range, centers around the village of Santo Domingo, which is 2 km northeast of the Real de Santa Eulalia (Fig. 1). REGIONAL GEOLOGY Thick-bedded middle Cretaceous limestones, probably Aurora, form the great bulk of the Cerro de Santa Eulalia. These beds, warped into a gentle dome, are severed on the west by a faultline scarp and truncated on the north by the Dolores Pass, a wind gap that

Jan 1, 1969

By H. C. Seeber

When considering all the aspects of a viable mine plan, it is important to estimate the accurate cost of mine ventilation, which varies as a function of airway dimensions, fan selection and ventilation network. Optimization of the overall elements of ventilation design can lead to very significant cost savings when capital and operating cost savings are considered in combination. This paper will describe a methodology, whereby ventilation costs can be minimized by optimizing the ventilation aspects of mine design and fan selection, by providing a brief description on ventilation theory and the factors which affect ventilation operating and capital costs.

Jan 1, 2002

By Cordner G. D

Titanium is not yet being produced on a commercial scale despite the facts that it has some impressive properties, and that it occurs very abundantly in easily-mined, highgrade oxide ores. One important reason for this is that the excellent mechanical properties of the metal were not widely recognised and appreciated until about twenty years ago.Another factor which has tended to retard progress in titanium metallurgy has been the realization that the oxide minerals of titanium are very stable; in other words, considerable energy has to be expended to reduce the oxide to metal. Table 1 presents the minimum energies which would have to be expended to produce one ton of each of the metals, aluminium, magnesium, titanium and iron from their respective oxides under theoretically ideal conditions at normal temperature and pressure. At higher temperatures, the minimum energy requirements are different, but the values for the various metals still fall in the same order as indicated in Table 1.In actual fact, the energy consumption in the case of any particular metal is higher than the ideal value by an amount which can be determined only by' the results of plant practice. The only general conclusion that can be drawn from the data in Table 1 is that it might eventually prove...

Jan 1, 1950

This paper describes the back analysis of the Pavoncelli tunnel, an hydraulic tunnel which was excavated between 1904 and 1915 in South of Italy and experienced severe damages during the 1980 Irpino-Lucano earthquake. A numerical simulation which is intended to study the behaviour of the tunnel during this seismic event is presented, also accounting for the swelling phenomenon that took place during and after construction.

May 1, 2009

By Alexandre Passos, Giorgio De Tomi, Tatiane Marin, Dennis Cremonese, Jacopo Seccatore

"In the mining industry, rock excavation is the first phase of the comminution process. Downstreamoperations such as secondary breaking, crushing and milling terminate the process, reducing the size of the grains down to the desired particle distribution. In such a cascade process, every process upstream heavily influences all the downstream operations. When rock excavation is carried out by drilling and blasting, a very economic source of energy (the explosive) is used at the utmost upstream; a more expensive source of energy (electricity) is employed for all the downstream phases. This means that managing the correct employ of the explosive energy is a form of managing the whole process, controlling the operational costs and therefore the profitability of the whole business. We simulated this cascade behavior, using as a starting point the experimental data obtained from small-scale block blasting, evaluating the influence of blasting on the grindability of the blasted material. This study showed the reduction in visible (particle size) and invisible fracturing (micro-cracking, resulting in a reduction of the Work Index WI). The particle size distribution for full-scale blasts was estimated by means of algorithm simulation, then we simulated the behavior of the comminution plant fed with the material possessing the characteristics previously estimated. All the costs (drilling, blasting, electric energy for the engines of the crushers and mill) were calculated and charted. The results show a clear curve that allows to estimate the minimum operating expenditure (OPEX) for the whole mine+plant framework. Being the WI dependent on the output on the blast, one can manage the performance and the expenditure of the whole comminution process by adjusting drilling and blasting parameters For any given mine, it is possible to perform a simulation and encounter the minimum OPEX scenario for the comminution process and managerial choices can be oriented accordingly. In a cascade productive chain as mining is, responsible strategic choices must be taken looking at the entire production framework. Compartmentalized managerial staff and budgets for separated production areas are not acceptable anymore. Somehow higher costs upstream can lead to effective decrease in costs and increase in efficiency and productivity of the processes downstream."

Jan 1, 2016

Retreating Cu-Mo Rougher Tailing by Flotation for Concentration of Copper and Molybdenum

Sep 13, 2010

By Thomas Deschamps

The main objective of this work was to investigate the effect of binder addition on the geochemical properties of surface paste tailings. Nine leaching columns were prepared following different disposal configurations: non-cemented, non-cemented with cemented layers, and cemented .The proportions of cement by total paste weight were between 0 and 1% for a whole column; the maximum amount of cement in a single layer was 2%. The material used was a sulphide-rich tailings coming from a Canadian hard rock polymetallic mine. The columns have been flushed once a week during 30 weeks. The recovered leachates were analyzed for pH, Eh, conductivity, and a wide range of elements. The drainage rates were monitored during the initial disposal as well as after each column flush. The results show that binder addition can improve the whole environmental performances for some of the studied disposal configurations. The disposal configuration consisting of intercalation of three 2% cement paste layers within non-cemented layers (0.66 wt.% of cement for the whole column), and the configuration with 2% cement added only in the two bottom layers (0,44 wt.% of cement for the whole column), showed the best results in terms of water quality and drainage rates.

May 1, 2007