Mining of Narrow Steeply Dipping Veins
Mining of Narrow Steeply Dipping Veins
Mining of Narrow Steeply Dipping Veins
Introduction
Many mineral deposits occur as steeply dipping narrow veins. In the past, these have been worked either by high cost labour intensive methods which are able to follow the vein with minimal dilution for high grade production, or by lower cost mechanized methods with large equipment, with increased dilution levels and lower production grades. A number of mines overseas are introducing mechanization into steep narrow vein mining, obtaining the increased productivity which this brings, but maintaining the advantage of high grade production associated with narrow stope widths. This paper reports on methods used in a number of mines in North America, Europe and South Africa, visited by the author in May and June, 1986. The study tour was sponsored by the Queensland Chamber of Mines through the Julius Kruttschnitt Travel Grant. The state of the art is described for metalliferous mines, ranging from labour intensive to mechanized operations. Also described are a number of techniques, used in steeply dipping coal seams, which could perhaps be modified for use in metalliferous mines. In the mechanized methods seen in coal mines, various coal cutters and shearers were in use. For adaptation to hard rock mining, these machines would have to be replaced by some other rock breaking method until advances are made in hard rock cutting technology. Suitable methods, depending on the type of rock encountered, could be roadheader type machines, impact rippers, or drilling and blasting. The latter may require further modifications to the method for prevention of blast damage to equipment installed in the face. The original intention of the author was to provide a detailed analysis of the labour, consumable stores, capital equipment and maintenance inputs for each method. As the tour progressed, it became obvious that most, if not all, of these inputs were dependent on conditions specific to each site, such as ground conditions, maintenance policies, and other "traditional" operating practices peculiar to the company, mining field or country. The author has therefore endeavoured to describe each mining method in sufficient detail to permit potential users to understand the principles of the method, and adapt it and specify the necessary inputs to meet the conditions of their particular sites. Mining methods are described under three headings: "Conventional" Labour Intensive Methods; Mechanized Methods; and Steep Narrow Coal Longwall Techniques. Some items of equipment not currently in use in Australia are then described. Finally, some concluding remarks on ground support practices and a summary of the state of the art are presented.
The units in this report are generally those used by the company concerned, although some conversions have been made by the author for consistency within sections or to correspond more closely with Australian usage. The unit for productivity used throughout is tonnes per manshift (t/ms).
1.
Principal Mining Consultant, AMC Consultants Pty Ltd, 8/135 Wickham Terrace, Brisbane QLD 4000 E-mail: bhall@amcconsultants.com.au
Figure 1 - ASARCO INC. - Galena unit narrow cut and fill mining
Mucking is by scraper with the winch mounted in the central raise timbering, or moved forward to a remote cribbed pass. For the first six sets up from the sill, the two outer compartments of the timbered raise are used as orepasses. Above this, the pass moves outside the raise and is cribbed up in the fill. This permits the scraper winch to be positioned outside the central ladderway/skipway compartment and provides a dead break in the pass a little above the chute. A 12" grizzly is installed at the top of each pass. After a heading is mined out, a 2-3 man fill crew takes 3-5 shifts for fill preparation and filling with mill tailings. Fill preparation involves cribbing the orepasses and remote access raises and passes, and sealing these with hessian for drainage and fill retention. Fill is generally to within 1'-2' of the back of the stope, leaving a gap for breaking the next lift. Because of the depth of mining - in excess of 5,000' - rock stresses are a major factor to be considered in mine planning. An extensive seismic monitoring network is installed for detection of seismic activity and rockburst prediction. Where recordings indicate a build up of activity at a stope, an attempt is made to induce the coming rockburst at blasting time. This is achieved by leaving the stope unfilled after a lift, and drilling out the whole back of the stope with 8' upholes, which are blasted with millisecond delay detonators over the full stope length. The sudden increase in stress in the pillar above the stope will frequently induce a "controlled" rockburst at blasting AMC Reference Library www.amcconsultants.com.au
time, when no men are in the vicinity. This method is used particularly where a stope is approaching the sill horizon of the stope above, with the crown pillar becoming smaller and more highly stressed. The 200,000 tons per annum is produced in 250 days by a total workforce of: Days pay: Underground Mill Maintenance Surface General Salaried Staff TOTAL
strike. These are used as ore passes from the face to the reef drive. As the face advances, laggings and packs are installed to extend the base of the panel of reef, and excess reef is drawn off down the diplines, which are covered over at the top, and at dead breaks at strike accesses, for maintaining safe access to and along the face. As the face advances, the "walls" of reef are reclaimed and replaced with waste walls (Figure 4). An "intertrack" is established between main levels to permit reef wall reclamation in the lower portion of the stope to commence sooner than would otherwise be possible. This prevents the pinching of broken reef by closure of the stope walls, and makes reclamation less labour intensive. Reef drawn from diplines above the intertrack is trammed by hand back to completed diplines through waste in the reclaimed area, where it gravitates to the chutes on the main level reef drives.
Figure 3 Durban Roodepoort Deep Ltd Reef drive and timber chute
trammed forward towards the face and tipped behind the triangle immediately below the intertrack. This method requires considerably more timber support in the back area than the wall shrinkage method. Until recent years, when pipesticks and other innovative systems have taken over many ground support applications from traditional timber packs in South African gold mines, the wall shrinkage method was preferred. With the advent of these cheaper and less labour intensive support methods, the triangular rolling panel method has become more popular.
Shrinkage Stoping
A fairly standard shrinkage method is used in the steeper reefs (Figure 6). The reef is divided into blocks approximately 60m long. A timbered access raise 1.8m square is maintained at each end of the block from the level below to give dual access to the working area and to provide a ventilation circuit. Ideally these raises are driven ahead of stoping to the level above, but can be carried blind with the stope. The latter case is not preferred, as it is then necessary to blast onto the top of the raise each lift. This requires entry via the other raise and ventilation of, and working in, a dead end until through ventilation can be reestablished. Outside the raises, the stope widens or narrows to the width of the reef, up to a maximum of about 3m, depending on stress and ground conditions (Figure 7).
The back of the stope is carried horizontal, and is drilled out using airleg mounted hand held drifters. Holes are 2.4m long at approximately 45 to the horizontal, giving an advance of about 1.8m. Reef is drawn out of the stope from the level below to maintain an adequate working height from the back to the broken rock floor. Extraction is by rail mounted compressed air rocker shovels loading into rail trucks through a series of drawpoints developed from a strike drive about 6m from the reef (Figure 7). The locations of the drawpoints are marked in the working area of the stope by paint lines extended up the walls as the face advances. Radio communication is maintained between the stope and haulage crews for withdrawing stope workers from above drawpoints to be mucked, and for controlling the amount of broken reef extracted.
Figure 6 Durban Roodepoort Deep Ltd Shrinkage stoping Typical longitudinal section Scale 1:500 (approx)
Figure 7 Durban Roodepoort Deep Ltd Shrinkage stoping Typical extraction level plan Scale 1:250 (approx)
upper section is divided into a skipway for materials transport, and a ladderway in which the air, water, fill and electrical services are also installed. Ventilation enters through the tubbing, travels along each face, and exhausts at the extremities, generally through a return air raise to the level above, but sometimes back down through steel rings in the fill to the level below. Broken coal is transported from the mining face to the central tubbing by a scraper conveyor on the floor of the stope. The head end drive assembly, powered by a 90kW motor, discharges above the pass compartment of the tubbing, and drives a single 26mm diameter chain down the centre of the conveyor. The conveyor is built up in 3m long sections, mounted on pontoons which float the conveyor to its new position during the filling operation. For a 3m length, the complete assembly weighs 1563kg, while the pontoons have a volume of 1680 litres.
In the stopes, fill retaining dams are built of timber and sealed with hessian, approximately 4m from the central access tubbing. This allows the tubbing to be carried below the level of the fill so that coal discharging from the conveyors can gravitate to the top of the tubbing pass compartment. The area between the dams and the tubbing is allowed to fill up with fill leakage, fine coal spillage, and general rubbish from the mining operations (See Figure 16). The seams are blocked out into mining areas 300m - 800m long on strike, and 210m or 110m high, between main levels. Manual mining takes place in smaller strike length blocks, by what is known as the "multiple attack" method (Figure 10). The stope is subdivided along strike into a number of "attacks", generally 30m long. The coal is drilled with horizontal holes and blasted down onto the conveyor. Timbering is installed as shown in Figure 11 for both hangingwall support and providing access to the top of the face. The back is supported with a reclaimable steel cap supported by the hangingwall timbers, which are butted onto timbers from the lift below to transmit vertical forces down onto previously installed support embedded in the fill. Faces are 3.5 - 5.0m high, 1.8 - 4.0m wide, and advance is 2.0 - 3.0m per blast. Typical labour organization is: 2 men drilling and timbering per attack 2 men blasting in the area 1 conveyor operator 1 organiser giving 10 men per shift for three attacks. Four overlapping eight-hour shifts per day are worked, consisting of three shifts with six hours at the face, and the fourth shift for such operations as timber and materials transport and tubbing extensions. A typical cycle is 20 days mining, followed by 5-7 days fill preparation and filling. In the manual method, fill preparation involves, apart from dam building and similar activities, the complete dismantling of the conveyor, lifting, and reassembling it above the timbering. Overall productivity is 8-10 tonnes per manshift.
Fill is produced by mining poorly consolidated sandstone in a surface quarry. The backfill preparation plant is located in old near surface mine workings. Feed from silos between the plant and surface passes through hammer mills, and +80mm material and trash are scalped off. Undersize material enters mixing tanks where water is added to achieve the desired density of 60% solids by weight. These tanks are tapped directly by boreholes leading to the fill distribution level 545m below surface. Pulp density is able to be simply controlled by maintaining the levels in the mixing tanks, water addition being increased or reduced as appropriate to maintain a balance between the viscosity, the head loss in the reticulation system, and the head created in the tanks. No pumps are used. From the fill distribution level the fill is reticulated to the coal producing areas between 686m and 1,250m below surface.
Fondon Colliery, Empresa Nacional Hulleras del Norte S.A. (HUNOSA) Asturias, Spain
The national Spanish coal mining company operates a number of pits in the Central Asturian Basin, mining coal seams in dips from 55 to vertical, at widths ranging from 0.45m to 2.0m. The narrowest areas are mined by timbered stopes, which are the most labour intensive of the operations. Although the actual cutting of coal is by shearers, this method is described in this section because of the high component of manual work in the mining cycle. Other mechanized methods are described in Section 4 below. Timbered stopes are used to mine coal seams as narrow at 0.35m, though the minimum mining width is 0.45m to permit adequate access. Mining takes place between main levels 85m apart, and faces are mined underhand, generally on a 60 plunge, giving a face length of 100m (Figure 12). The bottom 5m of the face is mined ahead of the main face, and timber supports and chutes are installed for loading broken coal into 3m3 rail trucks. Access is not normally available to the face from below. Because of the steepness of the stope, and the danger of falling materials, the number of concurrent operations permissible in the stope is severely restricted. To overcome this problem, the working day is broken into four shifts, with each shift carrying out a separate operation, viz: coal cutting timber transport into the stope
The distribution of labour between the various shifts depends on such factors as the stope width and ground conditions. Coal cutting is by shearer, which is hauled up the face by cables from a winch on the top access level. Two types of shearer are used: 1. 2. an HUNOSA developed machine, designated H1, cutting to 0.8m and a Russian Poisk compressed air operated unit, cutting to 0.45m.
Timber is lowered into the stope by small compressed air winches, and is stored on existing timbering at a number of elevations within the stope. Stulls of 75-100mm diameter are placed between hangingwall and footwall in lines parallel to the face, with caps and sills of half round laggings, approximately 3m long, over three or four stulls (Figures 12 and 13). The spacing of both rows, and stulls within a row, is 1.0 -1.2m. Stulls are sawn to size on site, and the ends of the stulls and notches in the laggings are shaped with hand axes. Laggings may be placed horizontally across several rows for working and safety platforms, although most of the work in, and travelling through, the stope is done on the stulls. The narrowness of the stope and the density of the stulls minimises the risk of men falling in the stope.
Figure 12 Fondon Colliery, Hunosa Manual steep seam method Long section
As the face advances, a fill fence is built on the fourth row of stalls back from the face at four row intervals (Figure 13). The fence consists of 10mm square wire mesh, supported against the stope walls by laggings. The stope is filled with a mixture of development waste and washery rejects, which is transported to the worked out area in sidetipping rail trucks, which permit direct tipping of the fill into the void (Figure 14). This method of filling only gives trouble if water inadvertently enters the fill
material, or if overbreak of the stope walls gives rise to irregularities, which are difficult to seal. On average, 30 men per day work in a timbered face to produce 145t of coal, an average of 4.8 t/ms. For this, 75t of fill is required in each stope each day. Transport of fill is not included in these productivity figures.
Figure 13 Fondon Colliery, Hunosa Section through manual timber stope looking down the face
Figure 14 - Fondon Colliery, Hunosa Cross section of top access to timber stope
Mechanized Methods
Vouters Colliery, Houilleres du Bassin de Lorraine Charbonnages de France, France
The general methods of working mechanised seams at Vouters are similar to those described for manual methods at this mine. The main difference is in the method of coal winning, which is by coal cutter in the mechanized areas. Two coal cutting machines are used in each stope, with the method of working being as shown in Figure 15. The workings on either side of the tubbing are phased such that the machine from one side can be retreated to the other side to permit fill preparation and filling on its side, without being delayed by, or interfering with, the other machine. A bridge is constructed across the tubbing, and the fill level is staggered by half the lift height on opposite sides of the tubbing. This is shown in Figure 16, which also shows the general arrangement of the tubbing and access crosscut. Three types of cutter are used, details of which are tabulated in the table 1. Availabilities of 90% are quoted for the machines during the mining phase, when only breakdown maintenance and "patch up" repairs are done. Preventative maintenance and more permanent repairs are done while the machine is withdrawn for filling, and on weekends. Ground support is similar to that for manual stopes. The coal cutters are equipped with arms for moving the steel caps from the brow at the bottom of the face to their new positions at the back of the face being mined. Stulls across the stope are not installed, so that the retreat of the machine is not impeded,
though occasionally, some may be placed to warn of stope wall convergence. These are removed before the machine is retreated at the end of the lift. The absence of timbering across the stope means that the scraper conveyor does not have to be dismantled before filling. It is merely loosened in the existing fill, and floats up as an entity as the new fill is placed. Whereas in the manual method the conveyor is located so as to minimize the amount of broken coal which does not gravitate onto it, in the mechanized stopes it is located towards the footwall to permit access for a sidetipping Eimco 630 loader, which cleans up along the hanging wall side (Figure 17).
Table 1
Manufacturer Type Length Mass Installed Power -cutting -movements Drum diameter/width Range of dips Mining widths -90 dip -65 dip Face height 2.45-4.90m 2.90-5.40m 3.0-5.5m 2.60-5.10m 3.10-5.10m 3.0-5.5m 2.10-5.60m 1.80-3.60m 5.5m 100kW 36kW 700/1800mm 55 -90 100kW 36kW 700/1800mm 55 -90 130kW 48kW 695/1700mm 50 -90 Ateliers du Nord de la France ANF-D10 13.0m 46-60t ANF-D11 13.0m 46-60t Sagem DRESSMATIC DIII 12.2m 35-50t
The typical labour complement for a mechanized face at Vouters is: 2 machine operators 1 man for ground support and clean up with Eimco 630 1 conveyor operator 1 organiser An additional 3-4 men circulate around a number of faces for fill preparation and stowing. Four shifts are worked, as in the manual method. A typical mining cycle is 25-30 days mining, followed by five days of fill preparation and filling. Productivities (including the fill crew) range from 10-40 t/ms, with an average of 17-18 t/ms.
The service raise contains a ladderway, separated from the rest of the raise by hoops connected by longitudinal steel straps, a skipway for materials transport:, and all service cables and pipes including vent duct, which carries air from forcing fans on the leve1 above to the mining faces. When mining of the faces is complete, a half height fill pour is placed in one face. The LHD is parked on this fill while a slot for the next lift is cut in the back at the service raise access, and the access itself stripped for the next lift. Mullock from this stripping is left on the floor, services are reconnected, the loader is driven up onto the new level at the service raise, and, after extending the orepass and manway rings and fill drainage system, the filling of both sides of the stope is completed. Fill is repulped deslimed mill tailings from the central processing plant for the La Crouzille division at Bessines, 12 km from Le Fraisse.
Surcharge loading is in the form of development mullock stowed on the slabs. Shear bars are 2.0m long and grouted 1.5m deep into the sidewalk . Reinforcing mesh is supported on these, and the tensile bars cut to the stope width are placed on the mesh. Premixed concrete is pumped from surface directly to the slab being cast underground. Circular holes are left in the slab at intervals for ventilation. In the trackless area of the mine, each cut is accessed by a ramp from the main access decline. Stope crews here are also responsible for trucking their production to surface in 7.8t trucks. In the shaft area, various combinations of "ramp-instope" and captive equipment methods are used.
direction until the end of the stope has been reached and an access has been mined to the second orepass. The direction of mining is reversed, the ramp is mined out, and the face is advanced to the other end of the stope block. After the slab has been cast the procedure is repeated. The slabs overlap at the ramp position so that nothing can fall from the top of the stope to the working area. The minimum distance between orepasses is therefore a function of the amount of overlap required, the lift height (4m), the gradient of the ramp (1:5), and the clearance from the edge of the slab to the ramp below (2.5m). Two accesses for men and materials are maintained through the holes in the slabs, one on each side of the stope between the ore pass and the stope limit. The two undercut and fill methods described can be combined in a variety of ways depending on the ratio of stope height to stope length. Where the stope length is reduced such that the vertical height cannot be covered by two stope length ramps as in the ramp-in-stope method, the top section can be mined by ramp-in-stope, the middle section by the captive method, and, when the lower level can be reached by a stope length ramp, by reverting to the ramp-in-stope method. Alternatively, if two orepasses are placed towards the ends of the block, the middle section of the stope could be worked similarly to the ramp-instope lower section method, but with the LHD captive on a ramp between the two orepasses, or the whole block could be mined by the ramp-in-stope method. Stope labour for all methods consists of two men on each of two shifts for a five day week and a third man on one shift for rail haulage of ore from the stope to the skip loading pocket in the shaft area of the mine. The overall cycle for one lift in a stope takes approximately one month. The stope labour productivity ranges from 7-25 t/ms and averages 15 t/ms. The overall productivity for the mechanized methods workforce is 12 t/ms. These figures are quoted for both cut and fill and undercut and fill, as the latter, as noted above, becomes more productive than the former as ground conditions deteriorate.
one on development. A stope crew carries out all operations in its stope. A typical cycle is: Mining (drill, blast, muck, ground support) Fill preparation Filling Ramp access stripping Total (average) 18-20 days 4-5 days 1-2 days 5 days 30 days
The figures in the following table have been calculated as the target production rates for a variety of stope widths and lengths for the two types of LHD. Stope labour productivities in tonnes per manshift (NB. 10 hour shift) can be obtained by dividing these figures by 4. All operations in the stope cycle as listed above are included. Actual productivities are approximately 80% of these target figures.
Figure 21 Dome Mines Ltd Sublevel blasthole stoping (after Robertson, 1981)
Productivity figures for some typical narrow cut and fill stopes and a modified Avoca stope are tabulated below for the time periods shown.
Modified Avoca (1stope 12months)
Cut and Fill (1stope (2 stopes - 3 12months) months) Manshifts Drilling Mucking Rockbolting Fill prepartion* Filling 56.5 86.3 71.9 35.1 79.9 159.3 489.0 1421 4397 9.0 37.9 46.0 35.9 10.6 23.4 95.1 248.9 898 2756 11.1
65.0 90.5 13.8 49.9 77.8 170.5 467.5 269 5542 11.8
* Includes fill fences, orepass extensions and drainage installation ** Includes scaling, manway maintenance and transport of gear
The chocks used by HBL are conventional longwall chocks with some modifications to account for the steep dip. They are linked together in groups of three (see Figure 23) to assist in preventing creep down dip, and are fitted with plates on the sides and rear to prevent ingress of fill into the working area. Chocks have been recently redesigned to extend the footings as well as the caps up to the coal face in order to counteract footwall failures experienced in the latest working areas. Hunosa's chocks are designed specifically for the steep dips. As well as the usual hydraulic legs across the seam, an additional jack is placed in the plane of the seam between the "sides" of the chock. This jack may be used during chock advance to push the overlying chock up dip, thus countering creep down dip. Since Hunosa uses a coarser fill than HBL, a total sealing off of the goaf is not required. The chocks accordingly are fitted with angled back plates, expandable with changing seam width, which overlap to prevent fill/caved waste rilling into the working area (See Figure 24). The method of manoeuvring the coal cutter along the face is also variable. HBL retains the conveyor pan, carried over from flat dip systems, on the steep dip chocks to guide the shearer. This is moved along the face by the "Dynatrac" system, whereby a drive on the shearer itself pulls it along a stationary chain supported on the conveyor pan. Cutting is done on the downstroke only. Hunosa uses a shearer mounted on a beam parallel to the face and attached to the chocks. The shearer is pulled up the face by a cable on a winch mounted in the top access drive. Cutting can therefore only upwards. Hunosa's chocks are fitted with handholds for providing access through the face. HBL fits its chocks with ladders between the legs, mesh between the ladders and face, and trapdoors on the ladderway at regular intervals. A wide range of productivities is quoted. Hunosa quotes 8.5 t/ms for one cut per day, with 20 workers in the face. HBL quotes previous achievements of 12-19 t/ms, and hopes to improve on this after making changes to the chocks. The crew, spread over four working shifts, consists of: 12 chock operators 8 shearer operators 5 chock maintenance men 2 shearer maintenance men 12 men outside the face, on tailgate conveyor etc.
HBL claims productivity improvements over the multiple attack method resulted in payback of capital costs for a longwall system in 15 months, corresponding to approximately 1 million tonnes of coal.
Downdip Longwalling
Hunosa are using a Russian system for mining down dip at the San Antonio Shaft (Figure 25). The panels being mined are 40-50m long, 1.5-2.0m wide and 100m high. Chocks are installed in groups of three; one main with two auxiliaries. One auxiliary in each group carries the hydraulic control panel for the group. Coal is cut by a continuous chain on which are mounted cutters, and the coal is swept along the face to the pass to the level below by the movement of the cutter chain along the face. The beam carrying the cutter chain is pushed forward by a ram on the main chocks. When a cut is completed, the auxiliary chocks are advanced individually, the main chocks remaining in position. All main chocks are then advanced simultaneously. The shields retaining the fill in the goaf are attached to and move down with these main chocks As the face is advanced downwards, a travelling way is established at one end of the face. The travelling way from the previous panel becomes the pass for broken coal to the level below.
At the commencement of mining the panel, a 5m layer of fill is placed on top of the chocks. For protection of the top access drive, a 7.5m layer of fill is placed on top of 2.5m high anhydrite packs. The chocks are initially installed immediately below the top access drive, and the top of the panel is supported with timbers until sufficient mining has taken place for the required packs and fill to be placed. Thereafter, the goaf is allowed to cave. At the bottom of the panel, a 5m pillar of coal is left for protection of the drive. This is recovered on the final retreat from the area. Mining of a panel takes approximately four months with a crew of three men per shift in the face. Installation of the system takes approximately one month. Space is made for a set of three chocks by moiling out the coal below the top access drive with a hand held pneumatic pick. The cutter chain is extended as each set of chocks is installed to sweep the moiled out coal to the pass at the opposite end of the panel. Installation of the equipment in a new panel is done by three shifts of four men, with a crew of similar size used for concurrent dismantling in the completed panel.
Figure 25 - Hunosa San Antonio Colliery Longwall advancing down dip Long section
The face labour productivity during the production phase is quoted at 13 t/ms. The overall productivity, including dismantling, reassembly, and placing of timber, packs and fill at the top of the panel, is about half this figure.
km/h maximum), the risk of the driver striking his head is low, and the potential severity of an injury so sustained is also low. Furthermore, the size and shape of a canopy truly able to protect the operator in a serious fall of ground is out of all proportion to the size of the machine. The manufacturer has therefore adopted the position that, since construction of an effectual, robust canopy is impractical, the operator should be able to escape from the machine as quickly as possible. To this end, the seat does not have a high back, so that the driver can retreat with minimal impediment. However, the driver's position has been ergonomically designed to maximize his comfort. Other concerns related to stability and cable reel protection. With respect to the former, one operator found the machine to be unstable, a concern not expressed by others. The manufacturer reports static overturning tests conducted with the machine jackknifed. Instability occurred on lateral gradients as follows. Bucket down Bucket up - empty - full - empty - full - 27.5 measured - 23.2 calculated - 15.8 measured - 11.4 calculated
With respect to cable reel protection, one operator had problems with muck from the floor packing around, and impeding, the cable reel and its operating mechanism, which are under the driver's seat. Fitting of a protection plate under the vehicle did not solve the problem. Both these problems appear to relate to operating practices at the mine in question, where floors were uneven, and a sharp transition from declined access ramp to horizontal stope floor permitted the tail of the vehicle to scrape the floor. Operators in Canada found it necessary to make modifications to the earth leakage protection of the trailing cable system to meet local legal requirements. The same may apply in Australia. The cable reel is designed to hold 85m of a light 25mm diameter cable on a drum with a 300mm minimum diameter. Compliance with Australian Standard AS 2802-1985 may have an adverse effect on cable reel capacity. Prospective Australian users would need to follow through these aspects with the relevant regulatory authorities. The author has not pursued these matters further. Availability figures were provided by two mines. Differences in operating practices, maintenance policies, and methods of calculating figures make comparisons difficult. However, from these raw figures, the ratio of operating hours: maintenance hours was 1.44:1 for the one case and 5.21:1 for the other. In the former case, operating conditions were difficult, all maintenance was done by maintenance crews and the machines were both operated and maintained by a contractor. In the latter case, operating conditions were good, some basic preventative maintenance was done by operators, and was not recorded as maintenance time, and the machines were operated and maintained by the owner. L'equipement Minier also manufactures a variety of drilling jumbos to complement the Microscoop, based on the Microscoop chassis. The "Minifore" is a single drilling boom which mounts on the bucket of the Microscoop, and can be supplied for either vertical or horizontal drilling. The fitting of the boom can be accomplished by the LHD operator without
27
assistance. An auxiliary hydraulic connection on Microscoop must be provided to power the drill boom.
the
The Microdrill CMM 500 HE and MAC 500 HE, mounting hydraulic and pneumatic drills respectively, are carried on a Microscoop chassis with the bucket replaced by the drilling boom and stabilizing jack. Some details and capabilities of the CMM 500 HE are shown in Figure 27. The MAC 500 HE has similar dimensions and capabilities, the only difference being in the drifter. These drills are comparatively new to the "Microgeneration" range, and no details from operating mines were available to the author. Dome Mines were preparing to take delivery of a MAC 500 HE at the time of the author's visit.
operation, the relevant fractures are stress induced, ahead of the advancing face. Broken rock is swept by the hammer onto a reciprocating flight conveyor installed along the face, which is 64m long. The present machine is operated by a 95:5 water:oil emulsion. It is planned that the next version, probably to be classed as the prototype production model, will be built to operate on 100% high pressure water. This is in line with the trend towards "hydropower" in the deep South African mines, where the pressure of service water in the mains from surface is used to operate machinery. The unit, weighing 5t, is moved by hydraulic rams along a toothed rack on the edge of the conveyor pan, on which the whole machine is mounted. Much of the current machine comprises the hydraulic powerpack, which would be eliminated by the move to 100% water power with the pressure derived from the head of water from surface. Productivities of the latest developmental machine have been very encouraging, being in the upper end of the range of productivities for South African gold mines. The potential for the system has been demonstrated by the achievement, in individual six hour shifts, of face advances between 5 and 6m, which is the industry average for a month.
minimal, random rockbolting, skewed into the sidewalls where necessary to permit installation, or with occassional random stulls installed above the height of the mobile machinery. No mechanized systems were observed where a closely spaced regular pattern of support was required in the sidewalls. Where closely spaced support was necessary in the narrowest stopes, stulls or packs were used in conjunction with scrapers or gravity for muck removal.
Conclusions
1. Technically feasible methods exist for mining steeply dipping narrow veins down to less than 1m wide, and could be applied to Australian deposits. Rockbreaking in metalliferous stopes currently appears to be exclusively by drill and blast methods, although other methods such as road headers and impact rippers may have application as technology develops. Drilling in stopes less than 2m wide appears to be the exclusive domain of handheld drilling equipment. The introduction of mechanized drilling rigs is in its infancy at these widths. Support of stope sidewalls may present problems in mechanizing operations in narrow stopes. However, where stope width and sidewall ground support requirements permit, there is a move to mechanization of the mucking cycle by the introduction of small electric LHD's capable of operating in stope widths as narrow as 1m. A number of narrow steep dip coal mining methods conceptually can be adapted to metalliferous operations, especially where the rock is amenable to non-explosive rock breaking techniques. The economic viability of any of a number of technically feasible methods will be highly dependent on such factors as local conditions, traditions, cost structures, and regulatory requirements and constraints.
thanked for allowing the author the time to conduct the study, and for support provided. The author is particularly grateful to the managements and staffs of the operations visited, for friendly, informative and comprehensive visits, for information readily made available, and for invaluable assistance in meeting the tight travel schedules involved. Thanks are also due for permission to copy drawings from company publications. Finally, thanks are due to the secretarial and drafting staff of Mount Isa Mines Limited for their assistance with the preparation of this report.
References
Cardenas, R C, 1983 Huaron Mines Reserves in Narrow and Winding Veins, World Mining, June 1983, pp36-43 Chadwick, J R, 1982 La Crouzille, Top Uranium Producer in France. World Mining, November 1982, pp40--43 Robertson, B E, 1981 Experimental Longhole Narrow Vein Mining at the Dome Mine. Paper presented at C.I.M. Underground Operators' Conference, Sudbury, Ontario, February 1981. Robertson, B E, 1984 Mechanized Narrow Vein Mining at Dome. Paper presented at C.I.M. Underground Operators' Conference, Bathurst, New Brunswick, February 1984. Reprinted in C.I.M. Bulletin, January 1986, pp69-44, and republished with updating as: Robertson, B E, 1986 The Mechanization of Narrow Vein Mining at the Dome Mine. Mining Magazine, October 1986, pp308-317
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Bibliography
Bolstad, D D, Hill, J R M and Karhnak, J M, 1983 U.S. Bureau of Mines Rock Bolting Research. Proceedings of the International Symposium on Rock Bolting, Abisko, Sweden, August-September 1983, pp313-320 Grazon, Gorlas, and Delaunay, 1983. Tranches descendantes sous dalles et tranches montantes remblayees dans la mine d'uranium de Bellezane. Industrie Minerale, February 1983 Heurley, P and Vervialle, J P, 1985 Le methodes d'exploitation du siege minier du Fraisse. Industrie Minerale, December 1985 White, L, 1986 Uranium Mining is Alive and Well in France. Engineering and Mining Journal, August 1986, pp26-30
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Acknowledgements
The author would like to thank the Queensland Chamber of Mines for the award of the 1986 Julius Kruttschnitt Travel Grant, which made the visits to the various operations described possible. The management of Mount Isa Mines Limited is