Search Documents

By Howard M. Graff

The Inland Steel Company operates five underground iron mines in the Lake Superior District. The two largest of these, from the standpoint of productive capacity, are the Sherwood Mine in Iron River and the Bristol Mine in Crystal Falls, both on the Menominee Range of Michigan. It is at these two mines that we have had our principal experience with belt conveyor installations underground, .At the Sherwood Mine, which has an annual capacity of about 450,000 tons, the first conveyor belt installationwas made in 1947 when conveyors were installed in the main haulage drifts on the 1200 ft, level, It was estimated at that time that approximately 5,000,000 tons would be mined from the stopes above the 1200 ft, level and. that if all this material could be moved to the shaft by belt conveyor, the entire cost of operation, including maintenance and the amortization of the equipment, would be about $,05 per ton, which was slightly more than half of what the cost would have been by conventional tramming methods employing locomotives and cars. Now, ten years later, the 1200 ft, level is almost completely mined out and we have actually removed. 5,200,000 tons and. the average cost of transporting this material to the shaft by belt conveyor has been $.095 per ton. Since there has been a substantial increase in hourly rate of pay and also in the cost of materials since 1947, we feel that the anticipated savings were very nearly realized., The 1200 ft, level installation at the Sherwood consisted of two 36" trunk line conveyors 815 feet long and 380 feet long, respectively, deliverying ore to the shaft. These in turn were fed. by three lateral 36" conveyors which paralleled the sides of the orebody, These gathering conveyors passed under draw points to which ore was brought from the stopes by mea S of scrapers or shaker conveyors, Shaker conveyors were used for a number of years in conjunction with the conveyor belt system of haulage but it was finally concluded that with this particular type of ore, the maintenance on shaker conveyors was excessive. Therefore, more recently, we have gone back to scrapers for the shorter distances and. are using belt conveyors in the scraper drifts for the longer distances. After this belt system had been in operation for about three years, it had proved to be so successful that it was decided to proceed with the same system or. a new level to be opened up 200 feet below. Instead of sinking the shaft and driving a horizontal drift to the orebody on the 1400 ft, level, an inclined conveyor drift was started near the shaft at the 1200 ft. level and driven downwards toward, the orebody at a slope of 15 degrees. The conveyor installed in this inclined drift is 36" wide, 990 feet long, is powered. by a 100 hp motor, and is capable of delivering 300 tons per hour to the 1200 ft, level shaft pocket. It is fed by lateral horizontal conveyors on the 1300 ft? 1400 ft., and 1425 ft, sub levels. Wien the inclined drift was started, the rock was scraped from the breast directly to the rock pocket at the shaft. As the distance increased., double scraping was employed, until the advance had reached 350 feet, At this point a portion of the final conveyor was installed, with a scraper slide near the breast, T e drift was then advanced be scraping on to the conveyor until a sufficient distance had been driven to permit the installation of an additional 200 feet of conveyor. This proved to be a very efficient and economical means of moving the development rock to the shaft and obviated the need of using cars or an inclined skip way, At the present time 85 percet of tie entire mine output is being mined below the 1200 ft, level and is delivered to the shaft via the inclined conveyor. The system appears to be working as well and as economically as did the conveyor systemonthe 1200 ft, level.

Jan 1, 1958

By James F. Kemp

Jan 1, 1914

By M. W. Airth

ON THE 6TH JULY in the year 1953 an announcement was made to the public that touched off one of the greatest staking rushes in the history of mining in Canada and resulted, in tl1e space of a few short years, in the development of the greatest reserves of uranium-bearing ore in the world. It was ?On this date that the famous "Blind River back-door staking bee" was completed with the acquisition of some 2,000 claims by the Preston East Do me-Technical Mine Consultant teams under the direction of Mr. Franc Joubin and Mr. Robert C. Hart. That story has been told from coast to Coast and is too well known now to re-quire elaboration. Suffice to say that, in the short elapsed time of less tl1an five years from that momentous summer of 1953, three of the Algoma mines are in full production, three more are rapidly nearing full production, five more 1are approaching the production stage, a new hard-surfaced highway, 108, has been built to serve the area, and a brand new, completely planned city, Elliot Lake, is being built to provide the amenities of living to the 22,000 people already in the area.

Jan 1, 1958

By D. A. Elkins

This report is the second of a series of three that will evaluate, in order, the metallothermic, carbothermic, and electrolytic methods of magnesium production. The historical development of the carbothermic process is outlined. Production costs varying from 38.9 to 21.8 cents per pound of magnesium are estimated for different versions of the process. Additional suggested process modifications are listed. The technical and economic potential of the different versions are compared. Areas deemed most promising for future research on the process are delineated.

Jan 1, 1967

"Chairman: Professor C. O. Frush Rapporteur: Mr A. C. SchmidtPapers:Planning of underground copper mining by J. K. Williams, L. Smith and M. P. WellsLinear programming models of copper leach and solvent extraction plants by W. P. C. Duyvesteyn, J. R. Cutland and J. R. L. RusselPlanning and forecasting stope production by D. W. Kappes and P. F. X. Moussett-JonesA proposed method of dynamic ore reserve assessment for a caving system of mining by A. F. Spinks and S. P. NicholsThe initial point made by Mr L. Smith in presenting his paper was that the planning aid outlined in the preprint was neither sophisticated nor complicated. However, it was a system which worked and was a very real way in which the computer was being used to assist miners, in particular, who at the present moment have very little experience in the use of computers.Mr Smith apologised for the fact that the editors of the paper had eliminated all reference to 'time'. This was very unfortunate because the essence of the system was that it saved time, which was one of the main points made in the paper.The operations being planned by this means cover the Rokana orebody which is a stratiform body mined by open stoping.The planner has to decide how much ore to mine from each base in each time period of his plan, that is, from three to four months hence to a year ahead. Normal procedure is for the planner to have discussions with the underground manager on some of the information he requires. The sort of information needed is (i) tramming limits for each face, (ii) extraction and grade factors, (iii) the present status of mining operations so that the planner knows where to start the new plan, and, (iv) details of any other mining consider¬ations that need to be incorporated in the plan.The objective at this point is to produce an initial plan for discussion purposes. All the information required comes from one sheet per stops and shows, inter alia, the stope identifier, the length along strike, the ore tonnage, the copper and cobalt grades and, finally, the extraction and grade factors.The program itself, among many other things, checks the data for validity, compatibility, completeness and credibility and prints error messages accordingly. Output from this system is in the form of a summary of production from each face, as well as giving the figures of ore reserve depletion. This summary indicates which slopes are to be mined in this time period, the original survey grades (copper and cobalt), the estimated production grades, how much is to be mined from each stope and the copper and cobalt production accordingly."

Jan 1, 2014

PURPOSE OF UPDATED ESTIMATE FOR MINERAL INVENTORY BLOCKS During the production stage of an open pit porphyry copper mine it was observed that the expected production grade, as determined from the block inventory estimates, often differed greatly from the head grade of ore delivered to the mill. It was determined that, for purposes of production scheduling and monthly forecasting, better in situ grade estimates for mining blocks were necessary. Because all of the bench blastholes were being sampled, and periodic holes were being drilled into the next underlying bench, much more sample data existed than was being used for estimating the grade of nearby mining blocks. It was reasoned that, by periodically updating the mineral inventory block file over the benches scheduled for mining in the next time period using all existing data, better estimates and forecasts for production grade could be made. BLASTHOLE SAMPLE DATA MANAGEMENT The collars of all blastholes in each mined bench were surveyed and assays run on the cuttings representing the full 12-m (40-ft) bench height. The blastholes range from about 5 to 6 m (16 to 19 ft) apart along the front of the bench as well as perpendicular to the front. Overbreakage often left larger spacings between successive blasts. Fig. 17-1 is a plan map of blastholes on a typical mine bench. All blasthole data were keypunched to a format resembling that of the 12-m (40-ft) composite assay data file as follows: [Collar Coordinates Compite Assoy Hole ID Northing East Elewotion Total Copper Oxide Copper] Because of the many blasthole samples available, and in order not to use excessive computer time for running kriged estimates for 30.5 X 30.5 X 12-m (100 X 100 X 40 ft) blocks adjacent to and beneath the mined area, the decision was made to average all the blastholes falling with each 15.2 X 15.2-m (50 X 50-ft) mined block, and use the mean value of the samples as a regionalized variable for purposes of assigning kriged estimates and estimation variance to adjacent unmined blocks. In other words, instead of using individual blasthole samples for making kriged estimates, the holes were grouped by blocks and assay values were averaged and assigned to the centroid of the holes within the block, which was then treated as a single regional variable for purposes of kriging. See Fig. 17-2. VARIOGRAM COMPUTATIONS AND KRlGlNG RESULTS With the many blasthole samples it was possible to compute directional and vertical experimental variograms for both sulfide and nonsulfide copper assays falling within the enriched mineral zone for the full 12-m (40-ft) sample support. Due to the close spaced drilling, excellent definition of the experimental variograms was possible, and the spherical model exhibited good fits. A three-dimensional kriging program was then run over the two or three mine benches involved in the inventory update, and estimated grades reassigned to all mining blocks falling within the range of the new blasthole assay data according to the anisotropisms of the deposit. Better confidence limits could then be assigned to scheduled mining blocks and better short- range forecasts made. An interactive kriging computer program was also applied for the purpose of determining the kriging variance or estimation of error for larger, irregular mining blocks representing the monthly production from a particular bench. The interactive program permitted the operator to enter the limits of the irregular block onto the screen of a cathode ray tube (CRT) as a series of points around the perimeter of individual gridded blocks making up the larger irregular block. The computer then was programmed to calculate the kriging variance of the larger block using all samples fall- ing within range. Thus the limits of estimated grade could be established at any confidence level. Fig. 17-3 illustrates the output from the interactive kriging program showing the sample points entering into the grade and kriging variance computations, and also the kriging coefficient assigned by the computer to each sample.

Jan 1, 1980

By Paul L. Cloke, S. R. Titley

Paul L. Cloke (Assistant Professor, Dept. of Geology and Mineralogy, The University of Michigan, Ann Arbor, Mich.) - On examination of the paper by S. R. Titley it appears that several errors have been made, many of them of a fundamental nature. It is the intent of this discussion to point out these mistakes, and, insofar as possible with the data at hand, to correct them. EXPERIMENTAL WORK Ferrimolybdite-Limonite-Water: The results of the experiments described are very difficult to interpret; nevertheless, a correct interpretation is essential as is pointed out subsequently. The difficulties become apparent when reasons are sought for the continuing change of pH after conversion of ferrimolybdite to ferric hydroxide and for the change of pH when -stirring is stopped after a steady value has been reached. In seeking an explanation of this experiment it seemed to this writer that the (temporarily) steady pH values might be due to kinetic rather than equilibrium effects. In this view the steady pH on stirring might be explained as a balance of several reactions proceeding simultaneously. The first of these would be the slow solution of ferrimolybdite,

Jan 1, 1964

By R. W. Brigstocke

History The birth of the mining industry of Canada took place about 1670. According to Salone (Salone E., La Colonisation de la Nouvelle France, page 205) the first discovery of ore in the country was made by an iron-master, Monsieur de la Potardiere, near Trois Rivieres, Quebec. Under French rule some prospecting was done; but at this time no one seemed to consider the working of mines practicable. It was not until the time that Begon became Intendant, when the Marquis of Vaudreuil was Governor, that mining was attempted. Begon was a patron of industrial arts and did his best to promote home manufactures. The mining industry was developed under his regime and discoveries of copper and lead ores were made in 1730. About this time the St. Maurice Forges were erected to treat bog-iron ore, and they were in operation for many years. Prior to 1747, iron pans to the value of £21,000 were manufactured, together with bar-iron and plough-shares. About 1769, that is, ten years after the conquest by Wolfe, stoves and other articles, and, later, potash kettles, appeared. What is still more interesting to us-or was, at least, to General Wolfe-is that cannon balls made at the St. Maurice Forges delayed his operations. A map dated 1744, a copy of which appears in the report of the Ontario Bureau of Mines for 1913 (Vol. XIX), indicates the extensiveness of the wanderings of the French and also shows the location of the lead mine on the east shore of Lake Timiskaming. Bog-iron ore was turned into metal, with charcoal for fuel, at Norman-dale, Norfolk county, Ontario, in 1813, and at Clementsport, Nova Scotia, in 1825. In a volume published in 1672, reference is made to the occurrence of coal on the Atlantic seaboard, and in 1677 Monsieur Duchesneau, the Intendant, levied a royalty of twenty sous per ton on all coal taken from Cape Breton.

Jan 1, 1932

By Joe L. Ratigan

INTRODUCTION Rock mechanics engineers are seldom concerned with obtaining the tensile or fracture strength of brittle rock at low mean stresses. The reason for this is two fold. Firstly, the behavior of many excavations in brittle rock is controlled by the inherent discontinuities in the rock mass. Because of the jointing and faulting, tensile strength is usually taken to be negligible. The second reason that tensile strength is seldom obtained in the laboratory is that the technology of utilizing this strength measure in design is undeveloped. Despite the fact that tensile strength of brittle rock is often ignored, there are a number of applications in rock mechanics wherein the knowledge of tensile strength is of fundamental importance. The apparent tensile strength must be known in a hydraulic fracture experiment if the state of in situ stress is to be determined from the initiation of the hydraulically induced fracture. In certain underground situations, the apparent tensile strength of intact rock beams defined by jointing or bedding planes is important in determining required rock bolting. Numerous other situations require a knowledge of apparent tensile strength including such high technology uses of underground space as geothermal energy extraction and LPG storage. Three observations are invariably made when intact rock samples are taken into the laboratory and tested to determine tensile strength. (1) The apparent tensile strength depends on the sample size (the 1 arger the specimen, the lower the tensile strength)

Jan 1, 1982

By R. J. Cornelius

Gold with minor associated copper mineralization in the area of what is now the Bougainville porphyry copper deposit has been known since the 1920s and was probably known in German colonial times, pre- 19 14, although no records persist other than a reference by Stanley (1922) to copper specimens from Kieta that he saw in Rabaul. Stanley did not at the time believe that they came from Kieta. Gold was first discovered at Kupei by Jack Comb in 1929 and the Kieta Goldfield was proclaimed in May 1930. In December 1932, the Tapu Reward Claim at Moroni - the site of the Bougainville Copper concentrator, was pegged by C.W.M. Evans who prospected up river from the West Coast. The Pumkuna (Panguna) Reward Claim on the site of the subsequent porphyry copper deposit was granted in February 1934 to G.A. Myers. n May 193 1, the Kupei Reward Claim was transferred to Bou- gainville Gold Options NL and later reverted to Jack Comb and then the Bougainville Syndicate, as did the Pumkuna and Moroni claims. Most of the interested parties had come from the Edie Creek rush on the New Guinea mainland. In 1935, N.H. Fisher, the geologist in Rabaul for the New Guinea administration who later became director of the Australian Bureau of Mineral Resources in Canberra, reported on the three gold areas. Norm Fisher described the gold and copper mineralization and its habit in great detail and this report had a major influence on the events leading to the discovery in the 1960s. Gold production up to the Japanese invasion probably did not exceed $50,000 in value. Half a ton of copper-gold ore from Kupei was sent to O.T. Lempriere's in Sydney for assay in the early 1930s.

Jan 1, 1991

By L. Nahai

In an exchange of views with the Program Chairman for the Technical Session on Industrial Minerals and Investment in Foreign Countries, it was agreed that it would be more interesting to approach the problem of mineral investment in foreign countries in a more general framework and not confine it to industrial minerals. Therefore, this Bureau of Mines report is proposed to examine current attitudes toward investments in foreign mineral resources and restraints that they impose on investments necessary to assure supply for future global mineral needs.

Jan 1, 1976

By B. M. Larsen, K. Heindlhofer

Jan 1, 1945

By E. W. Davis

THIS nation has been depending upon the Lake Superior iron ranges for most of its iron ore requirements for over half a century. Furthermore, it can continue to draw the major portion of its ore requirements from the same source for another century or more, but to accomplish this, great changes must take place in the present ore mining industry. Preliminary preparations for these changes are now underway and in the next few years decisions of far-reaching importance must be made. But to understand the situation, it is first necessary to know something of the nature and occurrence of the Lake Superior iron formation. So far as is known, the merchantable grades of iron ore remaining in the various Lake Superior ranges are as shown in Table 1, which shows that

Jan 1, 1947

By Diego Marquina, Jorge Terrones, Jose Morales

Si bien los incendios en el sector minero representan apenas el 1 % del total en todos los sectores, sus consecuencias son potencialmente críticas, por lo que resulta necesaria una evaluación constante de riesgos de incendio. En una evaluación de riesgo de incendio convencional (FRA, por sus siglas en inglés) se detallan los escenarios de riesgo de incendio a los cuales está expuesta una unidad minera, considerando las condiciones de la instalación, el análisis de los procesos, identificación de actos y condiciones subestándar, equipos u operaciones, reportes de incendio o amagos, entre otros factores que permitan identificar y analizar los escenarios de riesgo de incendio para definir su nivel de criticidad y aceptación del riesgo. Este enfoque diferenciado de FRA busca priorizar la inversión en protección contra incendios, al tomar en cuenta una tolerancia específica al riesgo de incendio, vida útil de la mina y la continuidad de procesos o recursos críticos para la operación en un nivel mínimo tolerable.

Sep 20, 2019

By James A. Browning

Following an unsuccessful attempt earlier in the program to develop and market a drill channelling machine, we have designed and tested a new continuous channelling method which reduces operating sound levels to below 100 db. This new method relies on a small diameter oxygen-enriched burner to produce narrow channels. Sound intensities are reduced substantially compared to conventional channel widths. They are further reduced by a water spray directed immediately above the flame. Baffles placed along the top of the channel will reduce sound levels to below 90 db. Cutting rates are greater than those of existing channelling burners with costs per unit of channel area substantially reduced for those cases where oxygen costs are reasonable.

Jan 1, 1981

By Pablo Hadzeriga

There are two places in the world where potash, trade name for potassium chloride, is economically recovered from naturally occuring brines by solar evaporation. One is at the Dead Sea and the other at Wendover, Utah. American Potash and Chemical at Trona, California recovers potash from natural brines, but uses fuel to accomplish evaporation. Under proper meteorological conditions the use of solar evaporation might be desirable from the economical standpoint. Of course, this may not be true when, if by artificial evaporation, a group of very valuable salts are recovered. This is the case of the American Potash operation at Trona. Solar evaporation has its limits. The working temperature is low. We may say, just to cover most areas of the world where solar evaporation may be practical, that it will take place between 18°C and perhaps around 45°C. In most cases the average working temperature expected will be around 25°C.

Jan 1, 1963

By E. H. Illsley

This bibliography is the first report of a Bureau of Mines research project on thermoelectric materials. Interest in such research has increased because of recent success in the design of commercial products, including generators and refrigerators, using the principles of thermoelectricity. This success made possible by the discovery of semiconductors that have much higher [effincies] than the metals and alloys formerly studied. Thermoelectricity is a term that includes a large number of phenomena relating heat to electrical energy. However, the moat important applications arise from two effects discovered by Seebeck in 1822 and by Peltier in 1834. When pieces of two unlike materials are joined at their two ends to form a continuous circuit, both of these effects may be observed. If the two junctions are maintained at different temperatures, an electromotive force known as the Seebeck voltage is generated. If, on the other hand, a current is made to flow through the circuit, the junctions of the materials will be found to differ in temperature. This phenomenon is called the Peltier effect.

Jan 1, 1959

By J. M. Provost

1. Estimates have been made for eight coal beds in Marion County, but only the Pittsburgh is important from a production standpoint at this time. The Pittsburgh bed has the largest recoverable reserve, and the Sewickley is second in importance. 2. Known measured and indicated reserves thickness of 14 inches and on 1,800 short tons estimated at 3,141 million tons, as of January lion tons is in beds 28 inches and more thick in all beds, based on a minimum per acre-foot of coal in place, are 1, 1955. Of this total, 2,983 mil-Areas in each bed were omitted from the estimate because available data relative to the bed characteristics were too meager for making an estimate that conforms with the definitions of measured and indicated coal adopted for this study. Should future drilling or development prove reserves in these areas, such reserves should be added to the total estimated reserves.

Jan 1, 1956

By Augusto Cabrera La Rosa

El presente trabajo revisa el panorama sobre el proceso de clasificación y explotación de los recursos minerales no metálicos en el Perú, con el objeto de promover el desarrollo de la minería no metálica. Los metales de origen mineral se clasifican en cuatro grupos: metales en estado nativo y minerales metálicos, combustibles minerales, aguas subterráneas y un gran número de sustancias minerales que no están comprendidas en ninguna de las categorías mencionadas, y que constituyen el grupo de los no metíllicos, llamados también minerales y rocas industriales. Asimismo, se muestra la clasificación establecida por el código de minería (combustibles minerales, minerales industriales, refractarios, material de construcción y minerales no metálicos diversos). Luego se describe el valor de la producción nacional de minerales no metálicos, como la situación actual de la minería no metálica en el Perú. Concluyendo que el Estado debe priorizar la prospección de minerales no metálicos de acorde a la importancia del desarrollo del país, asi como estimular la iniciativa privada para el descubrimiento de nuevos depósitos de dichas sustancias comercialmente explotables.

Aug 20, 1963

By BUREAU OF MINES

Mercury has been used by mankind since prehistoric times. It has been, and is now, widely used in industry and medicine and has unique properties which make direct substitution difficult or impossible. At present its prin- cipal uses are for electrical apparatus, control instruments, electrolytic soda preparation, mildew-proofing paints, pesticides, and medical and dental preparations. Deposits of mercury are less common than those of other base and precious metals, but some of the deposits are very large and have been ample to meet the world's past demands. The price of mercury has had many sharp fluctua- tions, but in general, the limited number of producers has resulted in a more stable price structure than that of many other commodities. The United States has been one of the main producers of mercury, and from 1850 to 1961 produced 3,247,704 flasks. Yearly production since World War II has varied from 4,535 to 38,067 flasks. An estimate of domestic resources of mercury producible at various price ranges was prepared for each mercury mine by combining specific knowledge of mercury mineralization with estimates of potential, based on geologic informa- tion and past production history. The sum of the individual estimates for each price range is shown in a table. A parallel estimate of nationwide production potential was made from a study of past price and production relationships. These were analyzed by plotting yearly production figures against the previous year's price, expressed in dollars of constant purchasing power. An equation was fitted to the data by mathematical means and was used to extend the curve into areas of higher than historic prices. Potential production rates at various price levels were taken from this graph. The estimated yearly production rates were used to make an estimate of total potential based on the premise that produc- tion rates and reserves have a proportionate relationship.

Jul 1, 1964