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By Dan Nilsson

INTRODUCTION There are many mine planning factors in sublevel caving, as in the other mining systems, which when varied can substantially alter mining costs and profit¬ability. In this chapter, the following four topics are addressed with regard to economic optimization of sub¬level caving: production planning, haulage level spacing, orepass spacing, and extraction cutoff. The costs used here are the right order of magnitude but since each mine is unique, they should be considered only as examples. The objective is to demonstrate the tech¬niques which can be used. In a real mine, the evalua¬tions must be done using actual costs and conditions. It is difficult to predict exactly the costs involved, and therefore it is valuable to perform a sensitivity analysis to evaluate the effect of making incorrect as¬sumptions. Since the mining industry is very capital intensive, the effect of the interest rate must also be studied. PRODUCTION PLAN FOR AN IRON ORE DEPOSIT Introduction The first and most important thing to do before de¬signing an underground mine is to establish a long-range production plan. Such a plan is necessary for all eco¬nomic evaluations and should provide information about the lifetime of the mine, how much ore and waste must be handled per year, how much development per year is required, etc. An example is given in the following section. Problem In an iron ore mine sublevel caving is used. The iron content is 42%, and the mine supplies a pelletizing plant with an annual capacity of 3 million t/a. The ore body is shown in Fig. 1. The length of the ore body is 1000 m and the width is 100 m. Each slice is 10 m high, and there are 10 m be¬tween crosscuts, each of which has an area of 20 m2. The density of the ore is 3.5 t/m3. The spacing between rings is 2 m, and the extraction is 100%. The iron con¬tent is 66% in the pellets and 6% in the tailings. The problem is to develop a detailed production plan. A typical sublevel caving sequence is shown in Fig. 2. Solution Amount of ore per meter of crosscut: 20 X 3.5 = 70 t. Number of crosscuts per slice= 100. Length of crosscuts per slice: 100 X 100 = 10 000. Ore from crosscuts per slice = 10 000 m X 20 m2 X 3.5 = 700 000 t. Number of sublevel caving rounds per slice: 10 000/2 = 5000. Area for each blast in sublevel caving: 10 X 10 - 20 = 80 m2. Amount of ore per blast: 80 m2 X 6 m X 3.5 = 560 t. Extraction: 100% (see extraction curve Fig. 12). Loaded ore per blast: 75% or 420 t. Loaded waste per blast: 25% or 140 t. Total: 560 t. Total amount of rock from each blast in the sublevel caving: Ore: 560 t of which 2 X 70 = 140 t from develop¬ment work and 420 t from sublevel caving. Waste: 140 t. Total: 700 t. The ore needed per year is 3 X (66-6)/42-6 5 mil¬lion t. Necessary number of blasts per year= 5,000,000/ 560 = 8929. Distance to develop per year = 8929 blasts per year X 2 m per blast 17 858 m. Total amount/year: Ore from development work 17 858 m x 70 t/m = 1.25 million t/a Ore from sublevel caving: 8929 blasts per year X 420 tons per blast = 3.75 million t/a 5.00 million t/a Waste rock dilution: 8929 blasts per year X 140 tons per blast 1.25 million t/a Total amount to hoist 6.25 million t/ a The amount of ore from development work will in¬crease a little if the horizontal drift is placed in the ore body and not in the footwall. In this example the ore loss in 20%, and the waste rock dilution is also 20%. After taking the ore loss into account, the lifetime for 100 m of the ore body will be:

Jan 1, 1982

By Leonidio Stegmiller

Since the petroleum crisis in 1973, charcoal1 is being considered one of the most important fuel for the Brazilian consumers, mainly for pellet's producers. In replacing fuel by charcoal mixed up with ore, obviously one of the first thing to be investigated is its influence in the product's quality. Special attention has been given to the metallurgical characteristics. This paper infocus the effect of increasing dosings of charcoal to the pellet's porosity and how other physical and metallurgical characteristics will be affected accordingly. This work was carried out using two different finesses of charcoal and dosing varied from 0.5 to 2.0%.

Jan 1, 1982

By J. Iannicelli

Spray drying is the principal drying method used in the production of over six million tons annually of beneficiated kaolin in Georgia. Thirty-three dryers in the Georgia Kaolin belt comprise the highest concentration of large tonnage spray dryers in the world. The steady growth of spray drying in the U.S. kaolin industry is paced by advances in design and development of large spray dryers which have led to large scale new applications in mining and related industries.

Jan 1, 1982

By Fumio Wada

Kuroko deposit is a stratabound massive sulphide and sulphate deposit associated with Miocene submarine volcanics and has been considered to be a prototype of meta- morphosed Palaeozoic and Precambrian equivalents. Since the discovery of Uchinotai deposit of the Kosaka mine in 1959, Dowa and other mining companies have conducted intensive exploration work in the Hokuroku district of northeastern Japan. Dowa has been placing its first priority on the exploration of Kuroko deposits and made a series of remarkable successes in finding the Kuroko deposits, i.e. Uchinotai, 1959; Matsumine, 1963; Uwamiki, 1966; Fukazawa, 1969 and Ezuri, 1975. It is emphasized that our successes have resulted from the proper application of exploration technologies accumulated through our experiences. Firstly we defined the Kuroko horizon and favourable felsic volcanism based on the detailed structural and volczno-stratigraphic study of the Hokuroku district, and then delineated the promising areas by means of the quantitative evaluation of geophysical, geochemical and mineralogical parameters such as the magnetic susceptibility and contents of alkaline and alkaline earth metals in the foot wall dacites and clay mineral assemblages of both hanging and foot wall pyroclastics and dacites. Practical application of these techniques is essential to our routine Kuroko exploration.

Jan 1, 1982

By Douglas W. Huber

During the past decade, a general awareness of the need for and subsequent development of increased safety in United States coal mines has taken place. Government and industry alike were stimulated into accelerating their actions through unacceptable accident statistics and costs (in terms of human misery dollar values). Those of us who work in the United States coal industry are aware of the development of governmental controls on nine safety. Both Federal and State regulations have placed increased responsibility upon mine owners and operators to insure a safe working environment and to provide training for miners. No other industrial group of workers has received such attention or, in our opinion, deserves it more.

Jan 1, 1982

By Warren D. MacEvoy

INTRODUCTION Good ventilation bulkheads (stoppings) can be obtained only by conscientiously adhering to four basic principles: carefully selecting the site for the bulkhead, adequately preparing the site, properly constructing and sealing the bulkhead, and checking the results with a smoke tube before leaving the area. If all these guidelines are followed during each ventilation project, a poor job will result only from the use of improper materials and cannot be blamed on carelessness or oversight. SITE SELECTION A perfectly constructed bulkhead in the wrong location can be completely worthless. Conceivably, a bulk-head in the wrong location could cause dangerous ventilation conditions that are difficult to identify and isolate. If the suggestions herein are followed, a considerable amount of work, time, and material may be saved. Perhaps the single most important step is always to select the exact site for a bulkhead before starting construction. Although the latitude in selecting a site may be limited by instructions or operating conditions, it is common to find poor bulkheads built within a few meters (feet) of a site that would have been far superior but was overlooked during the planning process. The ideal site for a bulkhead should have as many of the following features and conditions as possible: 1) To provide a solid bulkhead, the selected site should be in firm and unbroken ground. 2) To minimize the amount of work and material required and to minimize the danger of later damage to the bulkhead, the site should have as small a cross section as possible. 3) The bulkhead site should not be encumbered by interferences such as pipes, wires, ditches, wire roof supports, trash, muck piles, etc. 4) The walls and timbers of the site should be free from oil, grease, or tar that would inhibit adhesion of the sealant used on the bulkhead. 5) Unless special precautions are taken, the site should not have water seepage, standing water on the floor, or water-carrying ditches. 6) To expedite construction of the bulkhead, the site should have reasonable access to transportation, supplies, communication facilities, and compressed-air lines. 7) The site should have a reasonably level floor, allowing direction of the bulkhead door swing to be reversed at a later time if so desired. In a steeply sloping location, it may be possible to open the door in one direction only. SITE PREPARATION Once a suitable site has been selected for the bulk¬head, it must be prepared properly. All site preparation work should be completed before starting any construction work on the stopping itself. Unfavorable site conditions can be identified during the course of thepreliminary site preparation, saving time, effort, and materials that otherwise might be wasted. The follow¬ing preparation steps help assure the construction of a good stopping: 1) If wire mesh has been used at the site, a strip 457 mm (18 in.) wide should be cut and removed from the walls and roof at the selected location. 2) The exposed strip of rock should be barred down thoroughly to provide a smooth surface. 3) All obstructing materials should be removed, including old timbers, pipes, rockbolts, wires, etc. If a conduit must cross the bulkhead area, it should be located or relocated away from the floor, walls, and roof to allow a 6.28-rad (360°) seal around the juncture between the conduit and the bulkhead structure. 4) All loose muck should be cleaned from the site and a trench about 152 mm (6 in.) deep should be dug from wall to wall in the floor. The trench must be wide enough to accommodate the entire bulkhead, including the posts. 5) The rock surfaces of the walls and roof should be cleaned with a wire brush to remove as much loose surface material as possible. Thorough cleaning promotes adhesion of the bulkhead sealant to the surfaces of the walls and roof, thus promoting an airtight seal. 6) It is quite difficult to seal landing mats that cross a bulkhead. If such a crossing cannot be avoided, the bulkhead should be placed between the end and the first hole in the steel so at least one side of the bulk-head can be sealed easily and completely. That may have to be done prior to actual construction if the end is located on the opposite face from the seal coating of the bulkhead. BULKHEAD CONSTRUCTION The three principal considerations in bulkhead con¬struction are the type of bulkhead, materials to be used, and construction method to be employed. For many projects, the bulkheads are specified by the requesting agency, with no latitude for independent choice of the type, materials, or construction method. In such cases, any deviations from the specifications, for any reason, must be approved in advance by the proper department or by the project senior ventilation engineer. Types and Materials When a choice of type, materials, or method is allowed, consideration should be given to factors such as cost, required useful life, proximity to blasting concus¬sions, availability of materials, direction of permissible air leakage, and degree of airtightness required. Other factors to be considered include the amount of time available, accessibility of transportation, potential for interference with operations or production, ambient water conditions, availability of connections to com¬pressed-air lines, etc. Despite the multitude of factors to be considered, most stoppings can be analyzed easily and the proper choices can be made without much difficulty. The four common classes of bulkheads utilized in

Jan 1, 1982

By James C. Irvine

Pea Ridge Iron Ore Co., previously Meramec Mining Co., a joint venture by Bethlehem Steel Corp. and St. Joe Minerals Corp., mines and pelletizes iron ore at the Pea Ridge mine. The Pea Ridge property, now wholly owned by St. Joe, is located near Sullivan, MO, about 112 km (70 miles) southwest of St. Louis. The ore body was delineated in the mid-1950s by St. Joe during a lead exploration program. The first test holes drilled on the Pea Ridge magnetic anomaly revealed the presence of a large magnetite deposit; further drill¬ing in 1956 and 1957 confirmed that the ore body was minable. Meramec Mining Co. was incorporated in 1957 and shaft sinking began late in the year. Pro¬duction commenced in April 1964. The ore body is overlain by about 396 m (1300 ft) of flat-bedded sediments. It is tabular, about 792 m (2600 ft) long, and up to 182 m (600 ft) thick. It dips about 1.39 rad (80°) and is of unknown depth. The ore is mainly high grade magnetite with small zones of specular hematite. The wall rock is a Precambrian rhyolite porphyry. The mine was initially started with five major levels on 45-m (150-ft) intervals. Crosscuts were driven across the ore body on 58-m (190-ft) centers. Banks of stopes were mined between the crosscuts by a modi¬fied shrinkage stoping method (Fig. 1). This was done by undercutting a 12 m (40 ft) wide by 45-m (150-ft) long block and blasting horizontal "lifts" drilled on 1.5-m (5-ft) intervals. A 20-m (65-ft) sill was left between levels which contained the slushing drift, fin¬gers, crown pillar, and adequate thickness for support. By 1971 a "lattice" of pillars had been left and mining had progressed to the point that an orderly pillar re¬covery program was necessary (Fig. 2). INITIAL PILLAR RECOVERY PROGRAM The program was started in the upper western por¬tion of the ore body, furthest from the shafts. Due to the fact that the ore body narrows in the western extremity, the stope orientation was changed to mini¬mize required development. This left pillars 18 to 24 m (60 to 80 ft) wide and up to 79 m (260 ft) in length. The area selected for the start of the program had been developed between the uppermost level at 419 m (1375 ft) below surface and the 510-m (1675-ft) level. Due to the 91-m (300-ft) difference in depth, sublevels were driven on 15-m (50-ft) intervals and the ore was taken by rather classical sublevel stoping methods (Fig. 3). This left a structure with accesses compatible with drilling with 114-mm (41/2-in.) bore drifters mounted on columns. Most holes were drilled 63.5 mm (21h in.) diam and reamed to 100 or 127 mm (4 or 5 in.). Practical depth capability was about 21 m (70 ft) for holes above horizontal and 13 to 15 m (45 to 50 ft) for holes below horizontal. The interdependence of these long pillars necessitated that several of them be blasted simultaneously, making large blasts the only practical approach. Due to the required length of such loading cam¬paigns, it was necessary to select an explosive which could stand in the hole in an underground mine for a 4 to 6 week period. After some consideration a pump¬ able water gel was selected for both uphole and down¬hole loading. This afforded a high velocity, high density explosive which could be handled in large quantities. Loading the downholes was a fairly straightforward process. The holes were lined with an 8 mil polybutylene plastic sleeve and water gel was pumped into them with a double diaphragm pump mounted on the bottom of a stainless steel tub with a 81-kg (180-1b) capacity. Potential leakage into cracked areas of the pillar was

Jan 1, 1982

By Kadri Dagdelen

Design of the ultimate pit limit to direct long range mine planning has been a celebrated problem in the mining industry, and in mining academia, since 1965. Although many papers have been presented on this topic, even today, there is not a single solution technique which is widely accepted and considered universally best. Many philosophical differences exist among the pit planners concerning the use of different methods and it appears that resistance is very strong towards accepting other techniques if you are a believer in one. Whatever the case might be, the pit parameterization technique for the ultimate pit limit design is the latest and takes a different approach than its predecessors. There are already a number of papers which discuss the concepts involved (Francois-Bongarcon, 1976; 1978; 1980; 1981); however, the aim of this paper is to fill in some of the gaps which are thought to be missing in the previous discussions and, in addition, to present the recent research done on the subject at the Centre de Geostatistique de l?Ecole de Mines de Paris in Fontainbleau, France. This paper is divided into two parts. The first part will discuss practical aspects and interpretations of pit parameterization concepts with- out using mathematical or theoretical arguments; the reader who is interested in the theory should refer to Francois-Bongarcon (1978). The second part will present an approach to obtain a complete double parameterization of reserves using an inexpensive (computer costs) single parameterization algorithm. A case study of this approach is also included.

Jan 1, 1982

By Jack V. Hill

A unique type of grinding mill liner has been developed by Waagner-Biro A.G. of Austria. This liner system, termed an angular spiral lining (.ESL), was, successfully tested by Magma Copper Company's San Manuel concentrator in both their 10 Ft x 10 Ft (3.05m x 3.05m) and 12 Ft, 6 In x 14 Ft (3.81m x 4.27m) grate discharge wet grinding ball mills. Magma Copper's savings in power and grinding balls justified conversion of more than half of their twenty-six ball mills to the ESL system.1,2 The ASL system was provided by the Capitol Castings Division of Midland Ross Corporation, the exclusive licensee for such conversions. Capitol Castings proposed the following as expected benefits from installation of an ASL system in an 18 Ft x 21 Ft (5.49m x 6.40m) overflow discharge wet grinding ball mill at the Pinto Valley Operation of Cities Service Company: "(a) Reduction of power consumption per ton of ore milled = 12%, (b) Increased throughput = 6%, (c) Reduction of ball consumption = 6%."

Jan 1, 1982

By Roger M. Buckeridge, Charles D. Reese, Silvio M. Martino, William T. Carey, Ansel H. Graham

FUNCTIONS OF RAIL HAULAGE SYSTEMS The principal application of rail haulage in under¬ground mining is in the movement of ore or coal pro¬duction from near the mining areas out to a processing plant. In some cases, such production haulage may end at a shaft pocket or slope belt conveyor transfer station, depending on the type of mine and the geo¬logical placement of the mineral being mined. Main-Line Haulage The basic reasons for choosing main-line rail haul¬age are high-tonnage production per shift and long haulage distances. The choice between rail and other systems, such as belt conveyors, is not easy and requires an integrated study of many factors: tonnage, distance, grade, product size, number of producing areas, ore reserve, safety, capital cost, operational and maintenance cost, and the mining system. Projecting a main-line rail haulage system for com¬parison with other systems requires a choice of many variables which are interrelated. The project must start with consideration of the mining system, tonnages per day, distances and grades, and loading and dumping conditions. Rail mine cars of different sizes and types are studied. High tonnage demands consideration of wider track gages. A suggested minimum for metal mines is 914.4 mm (36 in.); for coal mines, 1066.8 mm (42 in.). The choice of the train size should be governed by tonnage per shift, physical factors of loading and dumping, cycle time, and the number of trains that can be operated and safely controlled. After the size of the train is determined, the type, size, and specifi¬cations of the locomotive can be selected. The loco¬motive may be trolley, battery, or diesel; selection guides are discussed elsewhere in this chapter. The advantages of rail main-line haulage are high tonnage capacity, flexibility, reliability, safety, low operational and maintenance cost, and dual usage of the system (supply haulage, for example). The dis¬advantages are high capital investment, high installation cost, the requirement for more maintenance facilities, and restriction to certain ore reserves and mine pro¬jections. Supply Haulage Supply haulage is a basic requirement of any mine, but the type and volume of supplies vary with the type of mineral being mined, the mining methods, and the size of the mine. A mine developed for rail main¬line haulage gains supply haulage through the addition of a secondary track system to extend the main line closer to the producing areas. Additional haulage equipment, generally smaller, is also added. It is neces¬sary to design into the rail system more switch bypass points and train control systems. The track system standards demanded by rail main-line haulage con¬tribute to the reliability and efficiency of the supply haulage function, whereas a supply track system to supplement other haulage systems often suffers from poorer track and installation standards. When belt conveyors are chosen for production haulage, a rail supply haulage system is often installed rather than using rubber-tired vehicles for supply haul¬age. The conditions of the seam or ore body create these items for consideration: haulage grade of more than 6% favors rubber-tired vehicles; longer haulage distances favor rail; difficulty in maintaining a smooth roadway favors rail; and movement beyond the end of a rail system generally requires rubber-tired vehicles, making a single system rather than a dual system attractive. Although operation of a supply haulage system may be concentrated on a single shift of a multiple-shift mining operation, proper planning and supervision are required if the haulage is to be efficient. Often the newer personnel at a mine are assigned to supply haul¬age which may add to the difficulty of obtaining maxi¬mum efficiency. Supply haulage by rail offers the advantages of high reliability, lower maintenance, and high efficiency over long distances. The disadvantages are less flexibility, high installation costs, and rubber-tired vehicles for use beyond the inby end of track often required. Personnel Movement The efficient and comfortable movement of the miner to and from his work area has a significant effect on his daily output. Minimizing man-trip time also becomes more and more important as hourly wages increase. Personnel are generally moved over a rail system by one of three methods: a specialized personnel carrier to which a certain number of men who work in a certain area are assigned; a combination vehicle, used to carry personnel but also capable of being used as a light supply/utility locomotive during a working shift; and man-trip cars towed by a supply locomotive. The first and second methods are used in coal mines where long distances make speed and comfort important. The third method is more common in metal mines. Equipment Movement As mining equipment technology has advanced, many mining machines and their subcomponents have grown in size. This trend requires careful planning for adding specialized carriers to handle the movement of larger equipment. Continuous miners, longwall equipment, raise borers, and ventilation equipment are examples of machines which may be difficult to move and which require special handling. The selection of the weight, speed, and operating characteristics of supply locomotives used to move equipment is important. Larger mines may often justify a special group of loco¬motives for this application.

Jan 1, 1982

By K. -A. Björkstedt

INTRODUCTION Strassa lies in the central part of Bergslagen, a tradi¬tional mining district, on the eastern side of the Stora Valley at an elevation of about 200 m above sea level. A railway siding runs between the mine and the Stora railway station from which there are railway connections to the shipping port and iron and steel works in Oxelosund, about 224 km away. The distance to the provin¬cial capital Orebro is about 60 km. The climate is typi¬cal for this part of central Sweden and is illustrated by the diagram of monthly precipitation and temperatures for the years 1968-1975 (Fig. 1). HISTORY There is no certain information as to when the Strassa mine was first worked, but it is known from sur¬viving accounts of mine inspectors that there were smelt¬ing works in operation in nearby villages in the 12th century. An example is the Gusselhytta ore smelting works, 10 km south of Strassa, which dates from this period. Around the year 1540 there were two smelting works in Strassa, the Upper Karberg and Lower Karberg works. Ore for these smelters was probably taken from Strassa and from the adjacent Blanka mine. In the year 1624 Strassa is mentioned by the painter Jons Nils Krook in an account of the iron mines in the Linde mining district (Linde Bergslags Jarngruvor). Several mines were listed in the area, the deepest being about 30 m. An impressive power installation is mentioned in 1639, including a piston system of lashed poles for transmit¬ting power from the Stora River to the Strassa fields. Its length was 2670 m. Common ground comprising about 20.2 km2 (5000 acres) of forest was allocated in 1689 for the furtherance of mining operations. Until the beginning of this century only the rich cen¬tral parts of the ore body were mined and these yielded, after handpicking, lump ore suitable for smelter feed. An example of the ore grades from these early times is an analysis of ore from the "Big Mine" (Storgruvan) from the year 1873: 48.5% Fe, 0.008% P, and 0.06% S. This same year a total of about 18 000 t was ex¬tracted from the Strassa mine. OWNERSHIP The mine was owned and run until 1874 by a min¬ing association made up of 119 so-called "bergsman," who were homesteaders often engaged in agriculture and timber-cutting as well. In that year the Strossa Grufvebolag (Mining Co.) was founded. In 1906 it was con¬verted into a joint stock company, the Strossa Gruveaktiebolag. This was acquired in 1907 by Metallurgiska AB for the implementation of Gustav Grondal's beneficiating and briquetting methods, for which the Strassa ore was well suited. The same year saw the completion of a new ore dressing plant with an annual production of 46 000 t of ore concentrate. In 1911 the mine passed to new hands, and in 1913 it was purchased by an Austrian company. Extensive new installations were made and in 1915 a new dressing and briquetting plant was completed with twice the capacity of the old one. In 1917 the Strassa mine was acquired by Granges. Be¬cause of unfavorable business trends and technical diffi¬culties, mining operations were brought to a close in 1923. Pumping kept the mine free of water until 1933 but it was completely filled ten years later. Up to 1950 the surface buildings and installations remained intact but the large dressing and briquetting plant burned to the ground in that year. Today only the machine shop re¬mains from this earlier period of operation, now housing parts of the Mineral Processing Laboratory. The decision to take up mining operations again was made in 1955 and construction work began the follow¬ing year. Of the old installation, only the "southern shaft" could be used for some development drifting after it had been completed with a new headframe. Other¬wise, all the buildings and installations required for the operations had to be rebuilt. New installations ready by 1960 were office and personnel facilities, a new shaft and headframe, a sorting and concentrating plant, a macadam plant, settling basins, pump stations, and a railway and yard with transport equipment. The instal¬lation was completed with two plants

Jan 1, 1982

By A. D. S. Gillies

The effectiveness of coal mine face ventilation and the ability of air flow to maintain satisfactory methane and respirable dust levels in the working face area has been studied through the use of scale modeling techniques and underground testing procedures. In particular, consideration has been given to examination of the effects of operation of a high air flow machine mounted dust scrubber unit on face air patterns. Results obtained through a cooperative study with the Peabody Coal Company demonstrate that new mining cycles can be considered which increase production efficiency while maintaining satisfactory ventilation conditions.

Jan 1, 1982

By John L. Halls

INTRODUCTION In this section, it is assumed that the reader is an engineer with sufficient training in underground mining work, preferably with budgetary control, to understand the mining methods outlined in other sections of this handbook. The mining-method descriptions are not repeated herein; only sufficient information is provided to enable the reader to comprehend the development of the costs presented. All costs are in 1977 US dollars, except where explicitly stated otherwise. CAPITAL INVESTMENT The capital investment in the underground mining portion of any mining property includes the following: 1) The development of original access involving adits, shafts, declines, etc. It includes the necessary equipment for transporting men, materials, and rock. 2) The purchase of mine equipment including loaders, underground haulage equipment, fans, pumps, electrical distribution, surface facilities associated with the underground mine, etc. 3) Sufficient preproduction mine development to assure steadily increasing production after startup. De- pending upon the mining method used, the preproduction development should provide a supply of three months to one year of ore that is developed and is ready to be stoped. Underground development also includes excavations other than shaft stations, loading pockets, sumps, etc., which usually are associated with the means of access and are excavated by the contractor. These excavations could include a crusher cavern, an ore- storage pocket, and underground workshops that may or may not be included by the shaft-sinking contractor. Portions of these could be delayed until after production begins, but it is preferable to complete them during the development phase, so there is no interference with the production buildup. 4) The staff and labor costs for all phases from inception through final production. 5) The provision of mine-surface equipment and facilities strictly applicable to the underground portion of the operation. The required facilities include a changehouse with a lamp room, a first-aid station, compressors, an electrical substation, workshops, a ware- house, water pipelines to the concentrator, water storage tanks, explosive storage, etc. A concrete-batch plant, a sand-fill plant, and boilers for heating may be required. 6) Other direct and indirect costs. 7) Engineering, design, and procurement costs. 8) Contingency allocations. 9) Contractor's fees, startup costs, and owner's costs. Original Access As part of a study for the US Bureau of Mines, Dravo Corp. conducted an extensive literature research survey covering underground operations in the United States between 1963 and 1973. Table 1 is compiled from the section of that study pertaining to the types of access to underground mines. Table 1 does not include distances attributable to the deepening of existing shafts, nor does it include shafts bored by raise drills or downhole machines that need existing underground workings for the disposal of the cuttings. During the period under review in the report ( 1963 through 1973). concrete-lined circular vertical shafts were the predominant means of providing the first access to a new underground mine; these shafts accounted for 80% of the original access in terms of the number of accesses completed, and 75% in terms of the total verti- cal linear distances. The inside diameters of the conventional circular shafts ranged from 3.7 to 7.0 m (12 to 23 ft). Most (72% ) of these conventional shafts were for production,

Jan 1, 1982

By Mehmet Guney

Acceptable working environments in production areas and service openings of Mindola Mine on the Zambian Copperbelt are currently maintained with an air volume of 1000 cu m/s and 5500 kW of installed refrigeration plant. It is planned to open up additional deeper areas of the mine for future production. Temperature problems, due to the increased heat generation from various underground sources are recognized as potential hinderances to efficient production. Findings of research work indicate that the present total heat output of 58,000 kW will inevitably increase by the second five-year period of the deepening program to 52,900 kW and by the fifth five-year period to 63,200 k\V. Estimated WB and DB temperatures, representing the total heat generation in stope faces and development headings during these same periods are predicted to increase to 53.8/40.2 °C and 35.5/42.1 °C respectively. It is assessed that the nominal refrigeration capacity needs to be progressively increased to 9,000 k\V This paper deals with the thermal problems in Mindola Mine, current heat output calculations, future prediction of temperature and heat load in various working areas of the mine with the application of computer programs and interpretation of data in order to find the effective solution for long-term planning of ventilation and air-conditioning.

Jan 1, 1982

By G. W. Johnston

INTRODUCTION For many years mining at the Stobie mine of the Inco Metals Co. consisted mainly of fill-type mining for higher grade ores and block caving or blasthole open stope mining for lower grade ores. Open stope mining, although quite favorable in the initial stoping phase, has always presented problems in the pillar recovery phase, especially at depth. In 1966, sublevel caving was introduced. This method is ideally suited for low-grade ore and can be applied very successfully to intermediate or higher grade ores providing proper controls are enforced. Sublevel caving has three distinct advantages: it is highly adaptable to mechanization; it recovers the ore in one operation; and it provides adequate ground sup¬port and safety regardless of mining depth. This method does, however, make extra demands, particularly for new types of supporting services. The constant presence of rock makes quality control of the grade of the ore being extracted essential. Increased de¬velopment, different types of ventilation, roadways, and maintenance are only a few of the many changes re¬quired. The care and maintenance of the fleets of diesel equipment and the modification of equipment is a con¬stant challenge. CHOICE OF SUBLEVEL CAVING The question arises, why mine by sublevel caving? Fig. 1 shows a typical problem encountered by the open stope or blasthole mining method. Near the surface there is little or no ground support problem. As mining progresses to an intermediate depth, however, ground failure starts to occur in the supporting pillars. If open mining is carried beyond this point, the critical depth is soon reached, accompanied by hanging wall caving, ex¬tensive pillar failure, and the danger of losing control of the mining cycle. The rock caved from the hanging wall mixes with and dilutes the ore being extracted from the blasthole stopes. Ore losses occur when pillar failure pre¬vents drilling and development of part of the ore. The interruption to the mining cycle causes serious curtail¬ment in production until ground failure is controlled and the mining method changed to adapt to the new conditions. Having recognized the limitations of open stope min¬ing, alternatives were sought. The first alternative (Fig. 2) was to fill the open stopes with sand fill be¬fore extensive pillar deterioration occurred, recover the pillar ore by fill mining, and then convert to total fill mining below. If the ore grade is high enough it can support the costly fill method. Recovery will be at a much slower rate and will therefore extend over a longer period of time. On the other hand, sublevel caving gives complete primary mining recovery in a short time and is very adaptable to mechanization, which is now a necessity to combat spiralling labor costs and give greater produc¬tivity per man. Sublevel caving is ideally suited as a mining method for low-grade ore bodies which eventually reach a criti¬cal depth for open stope mining. DEVELOPMENT The development phase comes first and of course its main purpose is to gain access to the ore and provide a route for ore removal from the mine. Initially a 20% inclined service ramp is driven in the footwall rock (Figs. 3, 4, and 5). It is preferable to start at the top of a mining block and drive the ramp down¬ grade so gravity can assist in mucking the blasted de¬velopment round, and also access can be gained more quickly to the initial production area. The ramp nor¬mally turns at 3.1 rad (180°) about every 152.4 m (500 ft) to keep it reasonably close to the rock disposal

Jan 1, 1982

By Daryl E. Radcliffe, Raymond M. Stateham

The Bureau of Mines, under a Cooperative Agreement with the Bear Coal Company, Somerset, Colorado, conducted a large scale evaluation of the effect of time lapse on roof stability. Time lapse in this study is the interval between exposure and permanent support of the mine roof. The study indicates that time lapse is not a critical factor in roof stability provided permanent support is installed before catastrophic failure begins. Concurrently with the time lapse study, investigations were made of stress distribution around resin-grouted bolts, the regional stress field, and anomalous geologic conditions. The behavior of the resin-grouted bolts, at least in the Bear mine, appears to fit the Coulomb-Navier Theory. Geologic conditions are primary factors in trends of roof stability that extend across the mine.

Jan 1, 1982

By E. M. Van Eeckhout

The National Uranium Resource Evaluation Program (NURE) has produced a mass of geochemical data in the form of data files on tapes. This paper presents a methodology for utilizing those tapes in a simple, practical manner. An example is given of an analysis of the Butte quadrangle, Montana, tape obtained from the Los Alamos National Laboratory (LANL). The example details the sequence of splitting the data file, plotting histograms, cumulative frequency curves, concentration plots, and methods for evaluation. All programs are written in FORTRAN. The case study was performed using a mini- computer (DEC PDP 11/70) and standard plotting subroutines.

Jan 1, 1982

By Louis A. Panek

INTRODUCTION Block caving is distinguished from other types of mining in its requirements that the mineral body to be extracted must cave after a relatively large tabular open¬ing is created underneath it, and that the caving must be predictable and controllable for both operating and safety reasons. A mineral body that caves readily may present special problems with respect to ground support in the mining access openings, where stability, rather than failure, is the desired objective. The complete ex¬traction of a caving block, which may measure 50 to 100 m on a side and 100 to 300 m in height, tends to create a large void, resulting in significant changes in the ground stress system and a migration of rock from all directions toward the void; these events begin very early in the life of the mine and lead to additional spe¬cial problems with respect to control of ground move¬ments in the mine working areas, in the rock adjacent to the caving stope, and on the subsiding surface above the mine. Thus, the mechanics of an undercut-cave operation involve a broad range of geotechnical subject matter, including the determination of the lithostatic earth stresses, the structural properties of a large jointed rock mass, the flow properties of the fragments after caving, the stability of tunnel-like openings, and the stability and displacements of the rock structure (1 to 2 km in size) that surrounds the ore body. The objectives are achieved by in-place measurement of rock properties, laboratory tests, theoretical structural analyses of the related prob¬lems, measurements of structural behavior in and above the mine, and comparison of theoretical predictions with actual performance in order to interpret and refine the predictive capability. Undercut-cave mining (Julin, et al., 1973) differs from most other methods in that the caving is an essen¬tial and inherent part of the excavation process and in¬volves the movements of millions of cubic meters of rock over distances of hundreds of meters, as compared to the gradual 2-m settling of the overburden typical of full¬caving longwall coal mining. Undercut-cave mining may involve a rock mass on the scale of a 1-km diam open¬pit mine, but mining is several times deeper than that in the open-pit mine; furthermore, a large portion of the rock mass is in a failed, rather than a stable, state. Ground-control problems are inevitable in a situa¬tion where ground movements are occurring on such a large scale, and where the rock mass is expected con¬veniently to cave at one location but to be stable at an¬other. To devise effective responses to these ground¬control problems, one needs to know first of all what is happening, not merely at the point where the symptoms are observed, but more broadly with respect to the be¬havior of the entire mine rock-mass structure over the total area of extraction, including the effects on the sur¬face as well as those underground. Monitoring structural behavior on this scale would require a sizable group of people, if they attempted to measure all the significant mining-induced changes of ground displacement and pressure, even on a 100-m grid of points. The practical approach is to conduct the mea¬surements activity by a judicious sampling, making in¬termittent, concentrated efforts in selected areas of lim¬ited extent to study problems of immediate importance; simultaneously a minimum level of surveillance is main¬tained over the total structure to delineate long-term be¬havior trends in the major structure, which provide the baseline reference data that are so essential for interpre¬tive purposes. Mine-structure behavior depends not only on the mine extraction configuration and sequence, but on the rock structural properties and how they change from one location to another in the mine. Increasing attention is being given to determinations of rock strength and de¬formation by performing mechanical properties tests in place and in the laboratory, and by drawing inferences from the mapping of geologic structure in mine openings as they are driven and from the logging of cores ob¬tained from exploration drilling that may be performed primarily for delineating ore reserves. Monitoring of mine-structure behavior and inter¬preting it in terms of the growing excavations and the structural properties of the rocks have significant im¬pacts on designing the major undercut-cave operations of undercutting and drawing the ore, supporting the ac¬cess openings, and controlling surface subsidence. The purpose of this chapter is to call attention to the com¬mon ground-control problems encountered in undercut¬cave mining, to relate them to the mining process, and to explain the geomechanical approaches to analyzing these problems with emphasis on methods that are most readily available to the mine staff. UNDERCUTTING, CAVING, AND DRAWING A distinguishing characteristic of an undercut-caving method of extraction is the objective that the rock mass fails of its own weight in a predictable and controllable manner, after it is undercut by excavating a tabular¬shaped void, horizontal or inclined. Induced or forced caving refers to procedures that may be resorted to when the ore will not self-cave over the desired span. Block caving, panel caving, and mass caving are terms used to denote the particular sequence of undercutting and ex¬tracting the total area of the ore body on a single min¬ing level. The discussion herein is not specialized to any of these variations; the word block as used does not imply anything as to sequence. CREATING THE UNDERCUT A common undercutting method is to drive a series of drifts and blast the intervening pillars. Theoretical structural analysis and practical experience show that ground-support problems in the undercutting drifts and in underlying access openings as well (the same princi¬ple applies even if undercutting is done by belling out drawpoints from the slusher drifts) are minimized by retreat-mining across the undercut slot according to a sequence such that the growing undercut area retains a

Jan 1, 1982

By Peter I. Brooker

The semi-variogram, the basic tool of geostatistical estimation procedures, must itself be estimated from the available data. From a numerically simulated two dimensional deposit, with known variability expressed through its semi-variogram, the standard arithmetic mean estimator is used to calculate experimental semi-variograms. These are determined on a series of data sets obtained from the deposit with various sampling grids which cover different areas of the deposit and which have a wide range of available data. The orders of discrepancies between estimated and theoretical semi-variograms are determined in this idealized situation; and this provides a basis against which to assess experimental semi- variograms made on field data.

Jan 1, 1982

By A. L. Mular

Se revisan brevemente varios procedimientos para el cálculo de coste de capital; no obstante, se discute en preferencia el método del factor estimado. Se proporciona una tabla de costes frente a los datos de parámetros de equipo, de donde los costes pueden ser adaptados a partir de los valores actuales de los índices de costes M&S para minería y molienda. Las ecuaciones de costes son utiles para preparar los programas de computador que realizan cálculos de coste de capital. Se ilustra la utilización del metodo del factor estimado.

Jan 1, 1982