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By Paavo Eerola

In May of 1984, it was decided that Outokumpu Oy would design and build a new nickel flotation mill to process ore from the newly discovered Enonkoski mine in eastern Finland. Since there hart peen a number of successful process charges in the existing mills, the design effort included the study and incorporation of several new developments, including all-autogenous grinding, "flash flotation", lane cell roughing, large cell cleaning and a computer-based, profit optimizing, process control system. The paper discusses the background and results of those studies, and a description of the resulting mill which will begin production in 1986.

Jan 1, 1985

By R. J. Hickson

The Grasberg copper-gold deposit was discovered in Spring 1988, just one year prior to the planned closure of Freeport Indonesia's original Ertsberg open pit copper mine. In spite of Grasberg's remote location, 600m above the Ertsberg, atop a hitherto inaccessible mountain, Freeport-McMoRan brought their latest open pit mine into production in the last week of 1989, less than two years after discovery. This paper describes the start-up of the operation, highlighting synergies drawn from both the original Ertsberg experiences and the current Freeport Indonesia underground mines.

Jan 1, 1991

By Ingemar Marklund

INTRODUCTION Today at LKAB, the primary mining method used is sublevel caving, independent of ore-body size, shape, dip, and geographical location. The mining of small ore bodies with relatively flat dip involves a lot of develop¬ment work, a large initial investment, big dilution, and large ore losses. In 1975 serious investigations into other possible methods, especially for those small ore bodies which do not lie in direct contact with other ore bodies, began. Different methods were studied and a number of in¬teresting ideas came from Russian mines. An analysis was made of one of the proposed methods, sublevel benching, in which mining is carried out from raises, and which uses the throw from the blast to place the broken ore. However, there were reservations about working in big open rooms at great depth, and doubts about the possibility of getting acceptable drilling results in an acceptable environment from raised platforms. Testing was therefore delayed for a time. However the investigations continued and in the fall of 1976, it was decided to try vein mining in a small ore body (Indian) (see Figs. 1 and 2). MINING RESULTS The results are now available from the mining of Indian and are as follows: Ore plan area 1200 m2 Level height 50 m Ore volume 60 000 m3 Ore tonnage 270 000 t Iron content 62% Dip 0.69 rad (40°) Ore thickness 14 m Ore compressive strength 26 MPa Ore C-factor (fracture density) 0.21 Hanging wall compressive strength 28 MPa Hanging wall C-factor 0.27 About 40 000 m of the holes were drilled by ring drills with an average productivity of 85 m/shift; loading of a round (2 rows, 750 m of hole) was usually accomplished in one shift. The total ore recovery was about 270 000 t with an average iron content of 59%, and three to four shifts were required per round (4700 t) to load. The principal of "throwing" the rock during blast¬ing worked well, and no ore remains on the footwall. ROCK CONTROL Before mining commenced at Indian, it was dis¬covered that the rock strength in its vicinity was rela¬tively poor. Since the ore is flatly dipping, and mining was to take place with open rooms, the strength of the hanging wall was of great importance. A supplementary program of exploration drilling was done during the fall of 1976 in order to assess fracture spacing and rock strength from the drill cores. Results showed that the strength of the hanging wall was low because of the many fractures. A decision was made to try cable bolting of the hanging wall by extending some of the ordinary blast¬holes. Cables (20 mm diam) were grouted in these larger holes and the rest of the hole was then utilized for ordinary loading (see Fig. 2). About 100 bolts altogether were installed and, to date, only small falls have occurred and the hanging wall is holding. Its exposed area is 2100 m2 DEVELOPMENT WORK On the 350-m level, four crosscuts (two per mining room), one longitudinal drift, and two elongated drifts (one per raise) for the raise climber were developed as shown in Fig. 1. Of the 230 m of drift, 60 m were in ore and the rest in waste. For mining room A alone, a total of 115 m of drift were needed with 30 m in ore. The drift dimensions were about 4.5 X 5.5 m. In the drifting on the 300-m level, two connecting drifts were driven from one existing transport drift forward to the respective raises. The total development was 93 m, all in waste (Fig. 1). The two raises having a length of 80 m each were driven at an angle of 0.68 rad (39°) from the hori¬zontal. They were driven along the footwall contact with a separation of 30 m and had an area of 10 m2. LONG-HOLE DRILLING The drilling of long holes with the ring drilling unit began in the fall of 1977 in raise A. The equipment consisted of an Alimak raise climber, equipped with a lower platform from which the charging was done.

Jan 1, 1982

By A. D. MacPhail

The Highland Valley Copper partnership was formed in July 1986 and was the logical result of a mine with a large ore reserve but no adequate mill, and a mine with a large, modem mill but dwindling ore reserves. The original partners -Cominco and Lornex -were joined in January 1988 by Highmont Mining Co. which contributed a two-line concentrator. During 1988/1989, the Highmont mill was moved 7.5 km to adjoin the Lornex mill, forming what is now called the Highland mill, with a current nominal capacity of 130 000 tonnes per day. The mine is located 220 km northeast of Vancouver, British Columbia and is close to four communities in which personnel make their home: Logan Lake, Merritt, Ashcroft, and Kamloops. The map in Figure 1 2.2 shows the general area. Kamloops is a major service centre and is serviced by two railways, two major roads and a regional airport.

Jan 1, 1993

By J. L. Huiatt

Kerosine, fuel oil No. 2, vapor oil, and fuel oil No. 6 were separated into general hydrocarbon fractions by ion exchange chromatography and adsorption chromatography. The collecting properties of these fractions were evaluated by flotation of molybdenum and copper minerals from a natural chalcopyrite ore. Paraffin fractions and aromatic fractions of low polarity produced the best molybdenum recoveries. Selective flotation of molybdenite from copper sulfide minerals was not possible with any of the chromatographic fractions.

Jan 1, 1972

By D. Blackwell, W. R. Reed, G. Ross, S. Peters, M. Shahan

Dust collectors for roof bolting machines generally use a dry box to collect the roof bolting material. Recently, an underground mining operation converted a dry box dust collector to a wet box dust collector with a unique exception from MSHA for testing purposes. Water is routed to the roof bolter from the main water line of the continuous miner. The wet box utilizes a water spray to wet the incoming material. Testing was conducted comparing the two different collector types. Respirable dust concentrations surrounding the roof bolter with the different collection boxes were similar. The main difference in respirable dust concentrations occurred when cleaning the dust boxes. The average respirable dust concentration during cleaning of the wet box was 0.475 mg/m3, and during the cleaning of the dry box the average respirable dust concentration was 1.188 mg/m3. The silica content of the roof material was high, ranging from 28.9% to 52.7% during this study. The results from this study indicate that using the wet box as a collector reduced exposure to respirable dust up to 60%when cleaning the collector boxes. INTRODUCTION Currently, the dry vacuum dust collection system is the standard dust collector on roof bolters in underground coal mining. It has been proven to be an effective dust control to prevent roof bolter operators’ overexposure to coal mine respirable silica dust during roof bolting operations [Potts eta l, 2011 and Colinet, 2013]. The dry collection system works as follows: drill cuttings are vacuumed into the precleaner. The precleaner dumps oversized material onto the mine floor while undersized material enters the dust box via vacuum airflow produced by a blower. The cuttings entering the dust box are either deposited on the floor of the dust box or into a dust collector bag. The vacuum airflow leaves the dust box through a final filter which removes any small particles from the airflow. Figure 1 shows the internal view of two dry collection system boxes; one that does not use the dust collector bags and another that uses the dust collector bags. Dust collectors that use dust collector bags to contain material inside the dust collector box have been shown to provide better protection to roof bolter operators than dust collectors that do not use the bags [Listak and Beck 2008]. The dry collectors are very efficient at reducing respirable coal mine dust exposure from roof bolting activities. Recently, Blue Mountain Energy’s Deserado underground coal mining operation, outside Rangely, CO, made a modification to the dry collector system to convert it to a wet collector system with a unique exception from the Mine Safety and Health Administration (MSHA) for testing purposes. The wet collector system still uses a vacuum system to pull material from the vertical roof bolt hole, but incorporates a water spray nozzle located inside the dust box to wet the incoming material. The flow through the nozzle ranges from 1.9–7.6 lpm (0.5–2.0 gpm) at 0.69 MPa (100 psi). The conversion to the wet collector system removes the precleaner and the cyclones inside the dust box. The wet material (sludge) then empties through a specialized valve in the bottom of the box. The final filter is still used to prevent respirable dust from leaving the dust box through the exhaust. During operation, the wet collector dust box is opened and hosed clean after roof bolting every place. After every two entries the final filter is changed out. Figure 2 shows the inside of the wet collector box.

Jan 1, 2019

By Paul Dean Proctor

The known northern ore zones of the New Lead Belt of Southeast Missouri extend to the drainage divide between the Meramec River draining north and east and that of the Black River draining southward. Intensive geochemical studies related to the environmental impact of the new mining area have been carried out in the vicinity of some of the newly developed mines, mills and smelter complexes (Bolter,1970; Gale, et al, 1973; Wixson, et al, 1969, 1977). Other investigations have related to fringe studies of the Northern New Lead Belt and the possible additions of heavy metals to the surface waters, stream sediments and selected aquatic life (Proctor, et al, 1975, 1976). Similar studies have been completed in the Joplin area in the Tri-State district, and in the fringe area near Springfield, Missouri (Proctor, et al, 1974). A surface geochemical study is currently underway on the lead-zinc-copper-cobalt-nickel province near Fredericktown, Missouri (Proctor and Sinha, 1977). This report summarizes some of the distribution patterns of heavy metals in the waters, stream sediments and selected aquatic life in the Upper Meramec River drainage basin just north and west of the known northern portions of the New Lead Belt of Southeast Missouri. The region studied includes the site of the proposed and somewhat controversial Meramec Dam and the reservoir site behind it. Some 3905 km2 (1508 mi2) are included within the drainage basin of the Upper Meramec River. Six distinct drainage areas are included within the drainage basin; 1) a combined, lightly industrialized and farm area around Rolla and Salem, Missouri representing the Dry Fork drainage; 2) a primitive, mostly forested area with some farms and cattle rancheseast and north of Salem, Missouri comprising the Upper Meramec River

Jan 1, 1977

By M. S. Klima

Tests were conducted to investigate the potential for recovering anthracite from preparation plant tailings. Samples were collected from tailings streams or slurry ponds of several anthracite cleaning plants. A washability analysis was conducted on each sample. The use of gravity-and surface-based cleaning processes for treating the 28x200 mesh tailings material was examined. Simulations were performed for various cyclone staging arrangements using published partition data for both water-only and dense-medium cyclones. Froth flotation testing was done on both precleaned anthracite and tailings material. Grade-yield curves were determined for various conditions.

Jan 1, 1995

By Lawrence K. Nordell

The selection of belt conveyor drive systems (*) is the subject of ongoing investigation. Frequently this choice has been governed by an incomplete perspective on the major considerations. A more comprehensive and objective approach is desirable. This paper presents additional factors which hopefully will broaden this perspective. BACKGROUND Conventional belt conveyors are powered by electric or electro-mechanical drives(*) mounted at the conveyor's head, tail, intermediate return span, or a combination of these (Fig. 1). A list of the drives reviewed is given in Table I. The following are main elements influencing the selection of the drive system: 1) dynamic characteristics, 2) economic factors such as capital cost, plant spares compatibility, reliability, availability, maintainability, and 3) environmental constraints. The principal dynamic characteristics of concern are the reactions of belt forces generated by the drive system during starting of the conveyor. Another attribute for consideration is the application of multiple driven pulleys used to 1) lower the conveyor's belt tension requirements, 2) increase the expected tonnage throughput, allowing for possible drive component failures, and 3) increase the plant spares flexibility related to the drive assembly.

Jan 1, 1977

By C. M. Goodall

Solids and liquid phase residence time distribution data is reported for a laboratory column flotation cell. The effects of various physical flotation parameters on residence time distributions are reported and analyzed. A plug flow model with recycle is proposed for the froth phase and a tanks in series model for the pulp. By comparison of two and three phase residence time -distribution data it has been shown that the use of liquid residence time distribution data to predict the residence time distribution of the solids in a column is not valid.

Jan 1, 1990

By G. L. Mowrey

A new coal interface detection (CID) method is under investigation by the U.S. Bureau of Mines. This novel concept is based on utilizing a passive infrared system (FIR) to measure temperature changes associated with the cutting of coal and adjacent strata. For this method to work successfully, a definite mechanical/hardness contrast must exist between coal and rock. Advantages include that FIR responds instantaneously at the coal-rock interface and it is minimally affected by dust and/or water sprays. Based on preliminary tests, results indicate that the PIR method can be applied successfully in mines having a hard immediate roof.

Jan 1, 1990

By F. H. Mertens

The costs of an open-pit operation can be classified according to the following hierarchy : -overall cost of production (per ton of ore) -costs of basic operations and functions -costs of auxiliary operations -costs of machines and equipment -maintenance costs. Production cost is a measure of the open-pit mine performance and should be the final target of the management control. It is also a measure of the share of the mine in the processing stages.

Jan 1, 1990

By Daniel A. Payne

For many years underground mining in the Sydney Coalfield has suffered from the effects of high horizontal stresses. It is only in the last 15 years that they have come to be recognized, understood, predicted and even avoided. This paper relates experiences in three coal mines in three different seams in the Sydney Coalfield. The effect of the direction of the principal horizontal stress on the direction of drivage of gate roads, access roadways and longwalls is described. Results of a monitoring program to measure the change of magnitude and direction of the principal horizontal stress in the gate roads as a result of longwall mining are also presented as well as the effect of the direction of principal horizontal stress on the shape of sandstone/gas outburst cavities.

Jan 1, 1998

By J. Marsden

Lift height, overall heap height and climate all play an important role in determining precious metals leaching rates, recoveries, operating costs and capital cost requirements for heap leaching operations. The effects of these factors are not easy to predict from testwork, and this can result in unexpected delays in metal production, lower recoveries than expected and higher capital and operating costs than necessary. This paper investigates the economic impact of lift height, overall heap height and climate on heap leaching and provides criteria to assist in the design and operation of leaching facilities. Descriptions of three heap leach operations in the western USA are provided to support the discussions and conclusions.

Jan 1, 1995

By R. J. Sweigard

Many states are considering the need to establish or modify policy relating to the disposal of coal ash. One of the primary concerns is the impact of leachate from the ash disposal sites upon groundwater and nearby bodies of surface water. A detailed site investigation was conducted at a coal ash disposal facility associated with a power generation station. The disposal site is situated alongside a major river. Monitoring wells were installed and soil samples were taken for characterization. Groundwater models were used to simulate migration of leachate from the disposal site toward the river. Such models could be used potentially to verify regulatory compliance. This paper addresses the validation and adaptability of these models to this case and their usefulness as possible tools to be used in future environmental audits.

Jan 1, 1992

By W. L. Freyberger

In many papers concerned with fundamentals of the froth flotation process, authors have considered the contents of a flotation cell to be comprised of two immiscible phases, the froth and the pulp. I (Of course, each of these phases is in turn made up of several solid phases plus air and water.) This generalized concept of the flotation system can be extended by considering the cell contents to include two phases: an extracting phase, the air bubbles plus the collector, which removes a selected mineral component from the second phase, the ore pulp. Thus, there is an analogy between a flotation system on the one hand, and chemical extracting systems, such a solvent extraction system or a leaching system, on the other hand. Such extracting systems may be operated in cocurrent or countercurrent fashion. This paper includes first a brief discussion on the general nature of cocurrent and countercurrent operation as applied to froth flotation, and then presents the experimental results of flotation of a specular hematite ore by each method, using a tall oil fatty acid as collector. Cocurrent Flotation A conventional flotation circuit involving roughing, scavenging, and cleaning flotation is illustrated schematically in Figure 1. In such a circuit the col¬lector, or extractant, and new feed ore enter together at the head of the rougher circuit, and flow cocurrently to the scavenger circuit where more collector is added. According to this practice, sufficient collector is added to the head of the rougher (and scavenger) so that the final tailings are of low enough grade to be discarded. Since the flow of collector and ore is cocurrent, the process consists of single stage contact between the extracting phase and the ore pulp. If the scavenger concentrate is recycled to the head of rougher, the overall circuit has some of the characteristics of countercurrent flow. In such a case part of the extractant (collector) is added at the discharge end of the circuit, and may

Jan 1, 1971

By Yi Luo

Subsidence events occurring above inactive room-and-pillar (R&P) mines, including abandoned mines or sealed portions of active mines, are classified as unplanned subsidence, because of their timing and distribution are unpredictable. When investigating such subsidence cases, many uncertainties exist and disputes often focused on whether the observed structural damages are caused by mine subsidence or other natural or artificial causes. However, solid evidence to identify the cause is often lacking. In order to conduct a complete and impartial investigation, thorough knowledge of subsidence theories and other ground control topics, as well as broad experience, are required.

Jan 1, 2011

By D. Scott Scovira, Karl Dufresne

Blasting in geothermally high temperature ground and/or extremely reactive ground is an increasing common challenge for surface and underground mines. Many current PETN based non-electric and electronic detonators have manufacturer published in-hole temperature limits ranging from 55°C to 65°C (131°F to 150°F). When confronting blasthole temperatures above the manufacturer published limits, operations may request site-specific guidance and qualification for use. The advised safe loading window for in-hole PETN based detonators is often too short to be practicable for production blasting. To overcome limitations of PETN based initiation systems, surface mines with high temperature ground and/or reactive ground are known to load blastholes with a qualified high temperature bulk emulsion that are top primed with a cartridge of temperature qualified detonator sensitive emulsion plus an RDX detonating cord downline. As there is no in-hole delay detonator in this system, blasthole timing is accomplished with surface delays. Although this system is safe, top priming and surface delaying does influence floor quality and top size fragmentation, and safe loading windows times reduce blast block size and increase blasting frequency. A mining company with a site characterized by geothermally high temperature ground plus highly reactive ground employed a blasting system like the one described above. The site desires to improve blasting performance by moving to an in-hole delay initiation system that can withstand the site ground conditions. At the request of the mining house, DynaEnergetics developed an electronic initiation system for application in geothermally high temperature ground and/or extremely reactive ground conditions. The DynaEnergetics IGNEO high temperature electronic initiation system is rated for in-hole use in ground up to 150°C (302°F) for 48 hours. The detonator consists of a high strength detonator with an HMX base charge and thermally resistant downline. To make up a primer, the detonator is inserted into a booster charge assembly that is fitted with a small HMX shape charge. This high energy primer is designed to be compatible and reliable for firing bulk emulsions, including those with high percentages of urea inhibition to combat reactive ground. Up to 1000 detonators may be programmed in 1 ms increments up to 1500 ms. The system is controlled by a digital firing panel that can be used with either a wired or wireless trigger. The system enables high temperature and/or reactive ground sites to utilize in-hole delay initiation and realize the associated performance benefits.

Feb 1, 2024

By Derek J. Barratt

The existence of a critical size buildup in an autogenous or SAG mill can be predicted from proper examination of drill core and ore samples. Methods of dealing with such critical size in a primary mill are reviewed in the context of ore properties, whether or not pilot plant-scale testwork has been conducted, and also with respect to the design of new plants or the retrofitting of existing operations. The implications of not planning for pebble crushing in situations where it is appropriate are discussed.

Jan 1, 1992

By A. W. Khair

Surface subsidence is a phenomenon mainly associated with longwall mining, although it can also occur if the coal pillar left to protect the surface deteriorates with time and results in the collapse of the room and pillar system. A method has been developed that utilizes physical mine models and laser holographic interferometry. This method is capable of measuring displacements as small as 5 micro- inches by means of generating displacement contours on a surface, even that of granular materials. Analysis of subsidence in room and pillar mining using models of various extraction ratios ad overburden depths along with two different types of overburden model material (sand and gelatin) has been made. The sand represented the extreme brittle conditions of geologic material while the gelatin represented visco-elastic behavior of geologic material. This paper deals with the analysis of data which has been gathered up to date, and mainly associated with the use of sand as an overburden model material. The analysis included the effect of variation of geometric parameters on surface subsidence. Among these parameters are: variation in seam depth and thickness, surface and seam inclination, and extraction ratio.

Jan 1, 1984