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A shortage of technical skills in all areas of the mining industry is being experienced worldwide. This has significant consequences for design, operations, productivity and safety and, therefore, needs to be seriously addressed, both in the short and longer term. It is therefore both a strategic and a tactical issue. Importantly, however, since mining is a long-term business, the shortage of skills must be dealt with from a strategic point of view to ensure the supply of the necessary skills for future mining. This paper will deal specifically with the provision of high level technical skills, that is, university graduates in mining engineering, and the outlook for the institutions that are currently producing such graduates.

Jan 1, 2009

By G. P. Conard, E. H. Rennhack

The generation of fibers composed of individual TiZn15 compound particles and their influence on crystallographic orientation in chill-cast, hot-rolled hypoentectic Zn-Ti alloys were explored. Metallo-graphic examination and X-ray diffraction data show that such fibers form a planar network after 50 pet reduction which eventually tends to suppress twinning at some critical reduction level (-90 pct) where the grain size approaches the interfiber, spacing. This leads to the development of a duplex-type projerred orientation through slip rotation consisting of an al-most ideal (0001)(2iiOj texture and a (0001) (1010) rotated component. Correlations between basal-plane rotation and secondary- creep rc11e sugg-est, however that prejerred orientation does not contribute in a major way to the creep resistance of these alloys. EFFORTS to improve the creep resistance of zinc through alloying without incurring a loss in fabricating properties have achieved only minor success with the exception of the very effective addition of titanium. Anderson, Boyle, and Ramsey,l for example, demonstrated that increasing the amount of thiselement from 0.05 to 0.23 wt pct produced a progressive, 45-fold increase in the inverse creep rate of hot-rolled

Jan 1, 1967

By F. V. Antonio, R. S. Pizarro

Masbate Gold Mines, dormant since WWI I, are now back in operation. The ore treatment plant incorporates carbon-in-pulp processing, utilized because of the ready availability of coconut shell charcoal in the Philippines and the high clay content of the ore. A detailed description of the plant, including the surprisingly fast and efficient alcohol stripping facility is presented. Operating data for the plant is provided together with an interesting performance evaluation of mechanical versus air- lift agitated leach tanks.

Jan 1, 1981

By Edmond Chin

The area encompassed by the political entities included in this study com-prises 11 million square miles or about 15 percent of the total Earth's land area. The bulk of the region is contained on the Asian continental land mass. Australia, the world's largest is-land or the world's smallest continent, alone is almost as large as the continental United States. The islands of New Zealand, New Guinea, Borneo, Sumatra, and Honshu are among the largest in the world. Collectively, they constitute a land area of 1.1 million square miles or an area about one-fourth the size of the United States. The region is noted for its great rivers. The Chang, Huang, Amur, and Mekong are respectively the fifth, sixth, eighth, and eleventh longest rivers in the world. The Ganges-Brahmaputra has created the world's largest delta, which covers 30,000 square miles. Abutting East Asia are the Sea of Japan, Dong Hai, Nan Hai, Bay of Ben-gal, and Arabian Sea. In the Southern Hemisphere, there is the Indian Ocean and the South Pacific Ocean. The Continental Shelf adjacent to the region constitutes a large proportion of the world's readily accessible shelf area. On the other hand, the Philippine and Java trenches and, further out, the Mariana Trench reach to formidable, abyssal depths, downward to 6,000 fathoms. Contrasting to this is Mount Everest, which reaches a lofty height of 29,028 feet. And then there is Xizang, whose high plateaus have been called the roof of the world.

Jan 1, 1992

By R. C. Thompson

The lead-zinc mill of Phelps Dodge Corporation, Copper Queen Branch, Mines Division, Bisbee, Arizona, is about 3 miles from the main hoisting shafts of the Junction and Campbell mines at Lowell. All the ore treated in the mill is obtained from these two mines. The mill was designed for an all-flotation treatment of 450 tons of lead-zinc ore per day, and has been operated at that capacity since the start of milling operations on Nov. 17, 1945. An expansion program is underway whereby the mill capacity will be doubled, but this article describes only the original program and presents results of the 450-ton operation. General Description The milling plant consists of an ore-receiving bin and three steel-constructed buildings housing, respectively, the primary crushing equipment, the secondary crushing equipment, and the ore-storage bins, grinding, flotation, and filtering equipment. The mill site is the one that was used for a previous copper-concentrating plant, which was operated over a period of years and then dismantled. There remained from this operation some milling equipment, a reservoir for mill-water supply, railroad tracks, tailings-disposal ponds, and a steel-constructed building, all of which were incorporated in the design of the present lead-zinc mill. The location plan of the plant layout is shown in Fig I. The flowsheet of the mill is based on the conventional scheme for lead-zinc treatment, however, several innovations were made necessary in order to utilize the existing building. Placing of equipment also was governed by the design of the building; thus making the mill unique in many respects. The following brief description of the mill building as adapted to the flowsheet points out the compactness of the installation and some of its unusual features. The first floor (Fig 2) houses the grinding equipment, with connecting conveyors for ore feeding, also an air compressor, two vacuum pumps, bins for lead and zinc concentrate storage with individual conveying systems for loading the two kinds of concentrates. Filtering of concentrates is done on the second floor, the concentrates discharging directly into the concentrate-storage bins. On this floor are the lead and zinc filters, two diaphragm pumps, sample-filtering equipment, sample-preparation room and a change room for the mill employees. The flotation machines are on the third floor, at elevations allowing a gravity flow of the lead and zinc concentrates to the thickeners. The annex to the building contains a freight elevator and provides space for all reagent mixing and feeding, also storage of reagents and grinding balls. All the water used for milling purposes is pumped from the Junction mine and

Jan 1, 1949

By R. C. Thompson

THE lead-zinc mill of Phelps Dodge Corporation, Copper Queen Branch, Mines Division, Bisbee, Arizona, is about 3 miles from the main hoisting shafts of the Junction and Campbell mines at Lowell. All the ore treated in the mill is obtained from these two mines. The mill was designed for an all-flotation treatment of 450 tons of lead-zinc ore per day, and has been operated at that capacity since the start of milling operations on Nov. 17, 1945. An expansion program is underway whereby the mill capacity will be doubled, but this article describes only the original program and presents results of the 45o-ton operation. GENERAL DESCRIPTION The milling plant consists of an ore-receiving bin and three steel-constructed buildings housing, respectively, the primary crushing equipment, the secondary crushing equipment, and the ore-storage bins, grinding, flotation, and filtering equipment. The mill site is the one that was used for a previous copper-concentrating plant, which was operated over a period of years and then dismantled. There remained from this operation some milling equipment, a reservoir for mill-water supply, railroad tracks, tailings-disposal ponds, and a steel-constructed building, all of which were incorporated in the design of the present lead-zinc mill. The location plan of the plant layout is shown in Fig I. The flowsheet of the mill is based on the conventional scheme for lead-zinc treatment, however, several innovations were made necessary in order to utilize the existing building. Placing of equipment also was governed by the design of the building; thus making the mill unique in many respects. The following brief description of the mill building as adapted to the flowsheet points out the compactness of the installation and some of its unusual features. The first floor (Fig 2) houses the grinding equipment, with connecting conveyors for ore feeding, also an air compressor, two vacuum pumps, bins for lead and zinc concentrate storage with individual conveying systems for loading the two kinds of concentrates. Filtering of concentrates is done on the second floor, the concentrates discharging directly into the concentrate-storage bins. On this floor are the lead and zinc filters, two diaphragm pumps, sample-filtering equipment, sample-preparation room and a change room for the mill employees. The flotation machines are on the third floor, at elevations allowing a gravity flow of the lead and zinc concentrates to the thickeners. The annex to the building contains a freight elevator and provides space for all reagent mixing and feeding, also storage of reagents and grinding balls. All the water used for milling purposes is pumped from the Junction mine and

Jan 1, 1947

By William F. White

It is a pleasure to have the opportunity of being here today to discuss the application of horizontal filters in the metallurgical industries. The use of horizontal type filters has been limited in metallurgical plants. However, in the recent past there have been new designs of this type equipment which greatly broaden their field of application. Through their use in many applications, important economies can be achieved over more conventional methods. The first of the modern continuous horizontal filters was initially developed by Oliver United Filters as their scroll type horizontal filter. This was an outgrowth of the old Oliver Sand Table that was limited to very coarse, free filtering pulps containing no appreciable amount of fine particles. The sand table was rather limited in application. The later continuous scroll type horizontal filter provided for a reverse air blow after discharge which reslurried the heel remaining and permitted the filtration of pulps containing finer particles. In the manufacture of phosphoric acid by the ?wet? process, the filtration of gypsum from the strong acid presented an extremely difficult filtration problem due to blinding of the filter media caused by crystallization of the soluble salts contained in the acid. A filter which allowed for continuous cloth washing was required, and this led to the development of the continuous pan type, and travelling belt type horizontal filters which provided for this need. Among the filters developed were the Prayon by Engineering and Industrial Corporation of Luxembomg for use in their phosphoric acid plants. They are one of the largest manufacturers of phosphoric acid in Europe. The Dorr-Oliver Traveling Pan Filter was also developed by Mario Giorgini, Manager of Dorr-Oliver?s Italian subsidiary, for use with the Dorr Company?s phosphoric acid process. Basically, there are four existing types of horizontal filters in commercial use at this time: The scroll discharge non-pan type, the continuous belt type (Lurgi), rotary pan type, and the traveling belt type. The construction and principle of operation of these various filters will be discussed.

Jan 1, 1958

Solid Energy New Zealand (SENZ) uses a 20+ year period for integrated planning of all its business activities. It has an ongoing program of coal resource assessment to optimise future mines within this 20 year period, using a six step resource and mine planning process. Desktop review and conceptual study at preliminary Levels 1 and 2 are based on general coal resource information. Further coal resource investigations are carried out for the more detailed evaluations in secondary assessment, prefeasibility study, feasibility study and detailed engineering from Levels 3 to 6. At Levels 1 and 2, completing conceptual planning, including preliminary resource evaluation and economic appraisal to the target cost uncertainty level of ¦33 per cent, is challenging in New ZealandÆs highly variable geological and geographical conditions. Where coal deposit geometries are relatively simple, approximate methods for determining pit limits using overburden strip ratios, seam thickness, physical boundaries, and quality cut-offs may be sufficient. Many New Zealand coalfields are however complicated structurally, with multiple seams and extensive folding and faulting. Coal seams are characterised by variable dip, thickness, and quality over relatively short distances. In the North IslandÆs Waikato coalfields, coal is typically overlain by weak sediments and clay rich strata that necessitate very flat cut and fill slope angles. The proximity of these deposits to major infrastructure (including towns, highways and rail lines), significant environmental features such as the Waikato river and lakes, and valuable dairy farmland, add further complexity. Traditional preliminary mine planning techniques applied in these conditions are insufficient. Pit optimisation software, widely used in the metalliferous mining industry since the 1980s, offers the ability to analyse many more specific factors affecting pit limits, mine layouts, and economics, with more rigour and for a much larger number of scenarios, than traditional methods allow. SENZ therefore decided to adapt and use æWhittleÆ pit optimisation software for Level 2 analysis of several Waikato opencast prospects. This paper describes this work.

Jan 1, 2005

By R. L. Pienaar

Paper written on project work carried out in partial fulfilment of B Eng (Min. Eng.) degree A study was conducted to investigated and compare the feasibility of two single level extraction layouts for an intact part of theorebody of the De Beers diamond mine, Koffiefontein. Two layouts were chosen and the tonnage profile, development schedule and production ring blast layouts were calculated by applying the information acquired through a literature study. Current principles used on the mine were also applied. The first layout chosen is described as a simple straightforward layout with 6 production tunnels spaced 15 m apart, and is referred to as Layout A. The second layout, named Layout B, is described as a complex chevron layout with 9 production tunnels spaced 15 m apart. This layout has a rim tunnel situated in the kimberlite ground, which is called a chevron and this made the layout more complex than Layout A. By comparing the financial results of the two layouts, it was found that Layout B is a more feasible option than Layout A. Sensitivity analyses show that both layouts are extremely sensitive to the revenue of the projects. Therefore it is recommended that Layout B can prove to be a source of additional tonnage to add to the total tonnage profile of the mine and to extend the life of mine by 20 months, only if certain financial/economic factors improve. There is, however, still some areas of the project that need to be investigated.

Jan 1, 2006

By Y. J. Wang

In the mathematical analysis of electrical networks, the electrical devices are often modeled using idealized network elements such as current sources, voltage sources, and resistors. An analogous approach is outlined here for the modeling of mine ventilation networks, which can be considered nonlinear resistive networks. The modeling technique proposed uses five basic (non-reducible) elements: the H element, the Q element, the D element, the R element, and the N element. These elements are characterized by their (a) pressure gain or loss, (b) air quantity, and (c) pressure-versus-quantity relationship, each of which may be known or unknown. In addition to the definition of the basic elements, the necessary conditions for which there exists a unique solution are discussed. This is performed using theoretical developments derived from electrical circuit theory and illustrated in example networks. The proposed procedures are useful in the solution and characterization of mine ventilation networks.

Jan 1, 1996

Acid mine drainage (AMD) forms when sulfide minerals are exposed to oxidizing conditions in coal and metal mining, highway construction, and other large-scale excavations. Iron sulfides common in coal regions are predominately pyrite and marcasite (FeS2), but other metals may be complexed with sulfides forming chalcopyrite (CuFeS2), covellite (CuS), galena (PbS), and sphalerite (ZnS). Pyrite commonly occurs with these other metal sulfides thereby causing AMD where Cu, Pb, and Zn are mined. Upon exposure to water and oxygen, sulfide minerals oxidize to form acidic, sulfate-rich drainage. Metal concentrations in AMD depend on the type and quantity of sulfide minerals present. The drainage quality emanating from underground mines or backfills of surface mines is dependent on the acid-producing (sulfide) and alkaline (carbonate) minerals contained in the disturbed rock. In general, sulfide-rich and carbonate-poor materials produce acidic drainage. In contrast, alkaline-rich materials, even with significant sulfide concentrations, often produce alkaline conditions in water. Approximately 20,000 km of streams and rivers in the United States are degraded by AMD. About 90% of the AMD reaching streams originates in abandoned surface and deep mines. Since no company or individual claims responsibility for reclaiming abandoned mine lands (AML), no treatment of the AMD occurs and continual contamination of surface and groundwater resources results. The oxidation of iron disulfides and the generation of acidity occur through several reactions (Stumm and Morgan, 1970). The chemical equations can be found in Caruccio and Geidel (2000). If any of the processes represented by the equations were slowed or altogether stopped, the generation of AMD would also slow or cease. The equations show that 1000 Mg of rock containing 1% sulfur, all as pyrite, would yield upon complete reaction 31.25 Mg of acidity. Therefore, 31.25 Mg of 100% CaCO3 would be required to neutralize the acidity. The alkaline content of overburden determines whether there is enough neutralization potential (NP) to counteract the acid produced from pyrite oxidation. Of the many types of alkaline compounds present in rocks, only carbonates and clays occur in sufficient quantity to effectively neutalize AMD. A balance between the acid-producing potential and neutralizing capacity of an overburden sample will indicate the ultimate acidity or alkalinity that might be expected in the material upon complete weathering. Accurate prediction and control of AMD before surface mining require an understanding of three important factors: 1) overburden geochemistry, 2) method and precision of overburden handling and placement in the backfill during reclamation, and 3) the postmining hydrology of the site.

Jan 1, 2002

By William Kent

In a paper by Mr. Frank Firmstone, published in vol. iv, of the Transactions of the Institute, on " Comparison of Results from Open-topped and Closed-topped Furnaces," the regularity of the average grade of pig iron procluced by the furnace under different conditions is expressed by means of a curve or diagram. The object of this note is to show that the graphic method, or that of representing variations by plotted curves or diagrams, is capable of a much more extended application to blast-furnace records, making a valuable auxiliary to the " furnace hook," and that by it the practical metallurgist may obtain at a glance information concerning the variations of the furnace during long periods, which could be found in the furnace book only after a tedious search. Plate XII is a graphic record of an anthracite blast furnace for one month. The table on page 554 gives the figures obtained from the furnace book, together with a record of temperature of the atmosphere, barometric pressure, and moisture in the air, which was obtained from a meteorological observatory not far from the furnace. In the records of a furnace, two of the variable qualities are final results, viz., the quantity of pig iron produced, and its grade; all the other variables are antecedent causes which produce the final results. Of these variable callses which contribute to the quantity of product and its grade, some are partially under the control of the furnace manager, others depend upon the state of the atmosphere, and still others are accidental. Of the first class are the character of the materials charged, and their relative proportions, their mechanical conditions, as wetor dry, large or small pieces, etc., the revolutions of the engine, the heat and pressure of blast, and the distribution of the charge. Of the second class are the number of pounds of air blown into the furnace for each stroke of the engine, the amount of moisture in the air, and the condition of the chimney draughts. These vary largely with the season of the year. Of the third class is the leaking of a tuyere or a temporary stoppage. To keep a record of this kind for a whole year, it is only necessary to procure a sheet of common profile or cross section paper, about 38 inches long and 8 or 10 inches wide, with cross lines ruled about 0.1 of an inch apart. The vertical lines, or those in direction of the

Jan 1, 1879

By William Kent

(Read at the Amenia Meeting, October, 1877.) IN a paper by Mr. Frank Firmstone, published in vol. iv, of the Transactions of the Institute, on "Comparison of Results from Open-topped and Closed-topped Furnaces," the regularity of the average grade of pig iron produced by the furnace under different conditions is expressed by means of a curve or diagram. The object of this note is to show that the graphic method, or that of representing variations by plotted curves or diagrams, is capable of a much more extended application to blast-furnace records, making a valuable auxiliary to the "furnace .book," and that by it the practical metallurgist may obtain at a glance information concerning the variations of the furnace during long periods, which could be found in the furnace book only after a tedious search. Plate XII is a graphic record of an anthracite blast furnace for one month. The table on page 554 gives the figures obtained from the furnace book, together with a record of temperature of the atmosphere, barometric pressure, and moisture in the air, which was obtained from a meteorological observatory not far from the furnace. In the records of a furnace, two of the variable qualities are final results, viz., the quantity of pig iron produced, and its grade; all the other variables are antecedent causes which produce the final results. Of these variable causes which contribute to the quantity of product and its grade, some are partially under the control of the furnace manager, others depend upon the state of the atmosphere, and still others are accidental. Of the first class are the character of the materials charged, and their relative proportions, their mechanical conditions, as wet or dry, large or small pieces, etc., the revolutions of the engine, the heat and pressure of blast, and the distribution of the charge. Of the second class are the number of pounds of air blown into the furnace for each stroke of the engine, the amount of moisture in the air, and the condition of the chimney draughts. These vary largely with the season of the year. Of the third. class is the leaking of a tuyere or a temporary stoppage. To keep a record of this kind for a whole year, it is only necessary to procure a sheet of common profile or cross section paper, about 38 inches long and 8 or 10 inches wide, with cross lines ruled about 0.1 of an inch apart. The vertical lines, or those in direction of the

Jan 1, 1878

By Carlos E. Roggero

The influence of high temperatures on the zinc roasting practice has been investigated by full-scale tests in fluid bed reactors operating at temperatures from 950° to 1150°C. It was definitely shown that, throughout the range, an increase in tem perature resulted in a 2 to 3 pct increase in zinc solubility because of decreased content of zinc fer-rite and lowered sulfide sulfur. The former is attributed to thermal decomposition of ZnFe,O,, and the latter to increased oxidation rates, ad longer retention times. Data on the gveatly increased volatilization of lead, cadmium, and silver above 1000°C, are presented. A considerable fraction of the iron present in the feed was also transferred at high temperatures from the coarse roaster discharge to the flue dusts. AS part of its metallurgical complex1-3 situated in a 12,000 it Andean valley at La Oroya, Peru, Cerro de Pasco Corp. (a wholly owned subsidiary of Cerro carp.) operates a 150 short ton per day electrolytic zinc refinery. Electrolysis is carried out at a cathode current density of 65 amp per sq ft with an electrolyte content of 55 g per liter of Zn and 150 g per liter of su lf uric acid. Due to the high iron content of the marmatitic concentrate used, a large fraction—about 15 pct—of the zinc reports in the leach residues. To minimize such losses, efforts have been exerted toward improving calcine solubility. One method, as noted in the literatre,, is to retard zinc ferrite formation by decreasing the roasting temperature. Unfortunately, such a procedure tends either to reduce the roaster throughput or increase the sulfide sulfur in the calcine. Consequently, this approach was discarded in favor of experimental work in the opposite direction, i.e., high temperature roasting. Full- scale tests were promising and resulted in the granting of a patent to Cerro de Pasco carp.' After a brief account of plant equipment and operating procedure, this paper describes the tests made and attempts to assess the results obtained. GENERAL PLANT LAYOUT AND OPERATION The Roasting Plant, see Fig. 1, comprises an agglomerating plant feeding three fluid bed roasters with a combined capacity of 10,000 tons per month. The Peruvian patent rights on this process were purchased from the New Jersey Zinc Co. The concentrate, assaying: is pelletized to a -3 +14 mesh product and fed to the fluid bed roasters. The roaster exhaust gases pass through a waste heat boiler, cyclones, a fan, and an electrostatic precipitator. Approximately half of the dust collected in the cyclones returns to the pellet-izing plant as one of the binding ingredients, the remainder joins the cooled bed discharge ('overflow') and is conveyed to the grinding plant. The -35 mesh product obtained, together with the dust collected in the electrostatic precipitator constitute the feed to the leaching plant. Batch leaches are made in vessels with mechanical and air agitation, the residues are separated and washed in Burt Filters, and pumped to storage ponds. THE FLUID BED ROASTERS Two reactors are as shown in Fig. 2, and the third,' of about the same configuration, is approximately 40 pct smaller. The total grate and freeboard (just below the gas outlet) areas measure 170 and 426 sq ft respectively. The feed, continuously weighed and controlled by Merrick Weighto-meters fitted with variable speed motors, enters the reactors at the top of the expanded bed level. The coarse calcine overflows through the opposite end by

Jan 1, 1963

By Y L. Ma, Y Yang, M Chen, W Bruckard

To improve the understanding of the reactions and mechanisms occurring during the leaching of sulfide ores at high temperatures a range of advanced characterisation and sensing tools have been combined to provide a sophisticated and powerful methodology to examine sulfide leaching systems. These include advanced electrochemical methods, surface chemistry and materials characterisation techniques such as synchrotron X-Ray Absorption Spectroscopy (XAS), X-Ray Photoelectron Spectroscopy (XPS), Scanning Electrochemical Microscopy (SECM), and in situ X-ray diffraction (XRD), together with new robust solid state chemical sensors for in situ real time monitoring of key aqueous chemical parameters. The methodology was employed to investigate various sulfide mineral leaching processes. For example, oxidation of arsenopyrite at elevated temperature involves complicated multiple reactions and the process is strongly influenced by the solution conditions, component speciation and mineralogy. The methodology was used to study this system both ex situ and in situ to gain a detailed understanding of the complex chemical reaction mechanisms operating. These mechanisms are discussed in terms of the observed leaching efficiency. The application of this methodology enables complex sulfide leaching systems to be understood more fully such that process efficiency can be improved.

Aug 24, 2022

By Ronald L. Wiegel

In the 2000 page Handbook of Mineral Dressing by A. F. Taggart, the subject of material or metallurgical balances is covered in only six pages of eve-straining type. Although Taggart's work was compiled before 1945, and there has been significant progress since then in most areas of mineral processing, the practicing metallurgist's approach to metallurgical balances remains exactly the same as described in those six pages. There are several portions of Taggart's text appropriate to introduce the subject of this paper which I would like to quote: "Milling Calculations. Computations of per¬formance in milling are complicated by the fact that it is a continuous rather than a batch operation, that the quantities of solid materials handled are large and usually mixed with water, hence difficult or impos¬sible to weigh. Fortunately it is possible to determine many facts concerning perfor¬mance without knowing weights, if the value of various constituents is known in some common unit, e.g., the content of some par¬ticular metal or mineral, or of water, or some ingredient soluble therein, or of par¬ticles of a particular size or falling within some particular size range, or the like."

Jan 1, 1979

By Hannweg L. A.

In a first for the diamond industry, Koffiefontein Mine in the Free State, South Africa, opted for the front cave mining method to extract a 90 metre high block of kimberlite ore. Block caving, because of its high initial capital requirements, resulting in a negative internal rate of return (IRR),was ruled out as a feasible mining method for Koffiefontein Mine. Front caving is essentially a combination of the traditional block cave and sub-level cave mining methods, but uses two or more levels of retreating drawpoints. Two extraction levels, in the case of Koffiefontein Mine, are spaced 12 m apart vertically, with the upper 478 m level (48Level) acting as the undercut/extraction level, and the lower 490 m level (49 Level) as the extraction level. Caving is achieved by undercutting a small column of ore, which is then drawn, on both the 48 and 49 Levels, until the planned extraction percentage is achieved. The next column of ore is then mined by advancing the undercut (48 Level), and the process is repeated until the entire ore block is depleted to the contact. Caving-by-gravity techniques of mining rely on closely controlled draw of the caved ore to ensure proper caving action and to minimize the dilution of ore with waste material from the open pit. This paper discusses the need for, and implementation of, a draw control system at Koffiefontein Mine — a computerized production management system, and a draw control system based on linear programming, to allow control over the draw rate and mining sequence — a major factor in all production planning to optimize production efficiency, ore recovery and cave management. Keywords: Draw control, Koffiefontein Mine, Front cave, Caving

Sep 1, 2001

By Ilmar J. Martens

The statistical information and cost data used in this report has been made available by NICKDATA INC. Since 1982, NICKDATA INC. of Toronto, Canada, has produced an annual comprehensive nickel industry cost study. Up to date information on 22 mines, 21 smelters, and 16 nickel refineries representing 30 nickel production operations in 17 countries are included in the computerized cost analysis. The report is compiled from information obtained through regular annual visits to major nickel operations and continuous monitoring of information on all nickel producers. The 1984 report which was completed this summer, reveals drastic changes in the nickel industry cost structures.

Jan 1, 1984

By J. P. Savely

This paper describes the development of a trim blast design to improve the stability of a final pit wall. Description is given of the effects of rock characteristics, rock type, and pit plans on the blast design. Currently, six benches have been mined with trim blasting. A more stable final wall has resulted and the successful program has reduced mining costs over an extended time period. Low-cost blast monitoring and field observation were used to determine peak particle velocity and blast damage. Using criteria developed from this relationship, trim blasts were designed and evaluated.

Jan 1, 1987

By W. W. Adams

Three hundred and twenty-eight mines and quarries, representing 36 States, took part in the twelfth yearly accident-prevention contest conducted by the United States Bureau of Mines and known as the National Safety Competition. Four underground mines and one open-cut mine (which represents the quarry and open-cut mine group), each the leader in its group, were awarded the ?Sentinels of Safety? trophy. Relative standing in the contest" was determined by the number of days lost from accidents in proportion to the total number of man-hours worked. Since 1925, when the National Safety Competition was inaugurated, an increasing interest has been taken in it by mining and quarrying companies as a means of stimulating mine and quarry safety by competition among similar plants. Each participating plant can, on a basis of its standing and record in the contest, determine to what extent its safety program is succeeding or to what extent it needs strengthening. With this information, each company that has a low standing in the contest can intensify its safety-promotion efforts the following year to obtain a lower rate and possibly win the trophy or achieve honorable mention in the competition. Except for 1929, the National Safety Competition of 1936 represented more man-hours of employment than any other year since the yearly contests began.

Jan 1, 1937