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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 H. C. Brunaugh, James Westfield, R. W. Whittaker

"INTRODUCTION The purpose of this circular is to describe the sealing off of a mine fire and subsequent use of carbon dioxide gas to extinguish it.A fire of electrical origin occurred in the Valier mine of the Valier Coal Co., Valier, Franklin County, I11., on March 29, 1949, at about 1:35 p.m. It was discovered about 3:35 p.m. in 5 south west intake airway inby 1 and 2 west panels. The No. 5 southwest airway had caved, and the fire was in the top coal beneath a large fall of rock. All men, except officials, were ordered from the mine. The mine officials decided to seal the fire, and four temporary seals were erected, which isolated approximately 200 feet of 5 south west entry. The seals were completed on March 30 at 2:45 a.m., but on March 31, at 3;40 a.m,, two of them were blown out by an explosion of gas that was produced by the fire. The two seals were rebuilt, the work being completed on March 31 at 12:00 noon, but they were blown out by another gas explosion at 12:10 a.m. on April 1. It was then decided to extinguish the fire with water if possible.To accomplish this, concrete-block dams were constructed. On April 1, about 4:25 a.m., when the dams were about half completed, water was introduced into the fire area through a 2-inch pipe.On April 4, J. G, Crawford, general manager of the Valier Coal Co., notified James Westfield, chief of the Vincennes (Ind.) Branch, U. S. Bureau of Mines, by telephone of the fire in Valier mine and requested assistance to combat it. Westfield, accompanied by F. J. Smith and H. C. Brumbaugh, Federal coal-mine inspectors, went underground and inspected the seals and dams. Later, during a conference with company officials, it was decided that the concrete-block dams were not strong enough to withstand the hydrostatic pressure that would be produced by flooding the fire area. It was then decided to use carbon dioxide gas to extinguish the fire. Four seals were completed, and 5 tons of carbon dioxide gas was introduced into the fire area, which extinguished the fire."

Apr 1, 1950

By D. Berberick, G. Sturgis, M. Board, B. Strickland

"Hecla Limited is nearing completion of the Lucky Friday Mine No.4 shaft located in Mullan, ID. The shaft is collared at the 4940 level of the mine and will have a final mine level depth of 8620, or 9,590 feet below the surface. After reaching the 7400 mine level the shaft encountered ground deformation higher then allowed for in the design for a circular shaft excavation and concrete lining. This additional ground deformation required portions of the shaft concrete liner to be reinforced with frameless, steel liner plate. To eliminate the installation of the liner plate from the remaining shaft development, the design of the shaft excavation starting at the 7650 mine level was changed to an elliptical cross section based on the overall stress field orientation as well as the orientation of the structure of the host rock surrounding the shaft. The elliptical cross section was instrumented to confirm the shaft excavation and lining was performing as designed. This design change has allowed the shaft to continue to the planned 8620 mine level without the need to install liner plate in the shaft or change the overall functionality of the shaft or the Lucky Friday life of mine plan.INTRODUCTION Hecla’s Lucky Friday Mine is a deep, underground producer of silver, lead and zinc in the Coeur d’Alene mining district of northern Idaho (Figure 1). Primary access into the mine is via the Silver Shaft, a 5.5-m (18-ft) diameter, concrete lined shaft sunk to an original depth of 1,890-m (6,200-ft). Current production is from the Gold Hunter vein system, approximately 1,525-m (5,000-ft) northwest of the Silver Shaft.Lucky Friday’s #4-Shaft is a 5.5-m (18-ft) finished diameter, internal shaft, or winze, being sunk to the north of the Gold Hunter vein system, from the mine’s 4760 level, 1,743-m (5,720-ft) below the surface, to the 8620 level, 2,923-m (9,590-ft) below the surface. The shaft is designed to provide access to the deep, high-grade zones of the Gold Hunter in order to increase the mine’s production and operational life. Design and engineering began in 2006, with construction starting in the fourth quarter of 2009."

Jan 1, 2016

By W. E. Anderson, T. M. Brady

As part of a larger study to identify factors influencing the practical operating life of wire rope used on large draglines in surface coal mining, field trips to operating surface coal mines were made during late 1975. Wire rope performance information was obtained from personnel at 13 mine sites west of the Appalachians. Seventeen pieces of equipment, utilizing bucket sizes ranging from 30 to 220 cu yd (23 to 168 m), were documented. Draglines at large US surface coal mines can be represented as 120 machines having 60-cu yd (46-m3) buckets, moving 1.8 X 109 cu yd (1.4 x 109 m3) of overburden during 6650 hours of productive availability per year. Based on the limited field study, the calculated downtime resulting from rope-related causes averages 105.6 hours per year for each typical 60-cu yd (43-m3) dragline. The total loss in production from this source alone is thereby estimated at about 25 x 106 cu yd (19 x 106 m3) annually. Although study findings did not reveal anything about dragline rope life that wasn't known or practiced by one operator or another, most operators can expect improved rope life by periodic review of their practices and by adopting one or more life-improvement measures employed at other sites. Since rope productivity was found to vary by a factor of five among the sites studied, considerable improved average performance is possible. As little as 10% overall improvement would increase national productivity by 2 million cu yd (1.5 X 106 m3) of solid overburden removal. Associated savings in rope costs would be some $2 million annually, plus a large reduction of labor-related expenses. This report outlines preliminary findings which may prove useful to many operators in extending wire rope life. Some of these modifications or adjustments in maintenance/operating practices would not require capital investment. During the field survey, no company was found which had tried cost reductions by using presocketed dump ropes, or dump rope sheaves that opened like a snatch block, so these are suggested as possible improvements requiring modest capital input. Data Acquisition During the field visits, mine operators provided rope life information about their machines. Some of the rope life data consisted of gross averages; other data was in the form of detailed records, including rope construction and manufacturer, cutoff and resocketing periods, end-for-end switching, and downtime for rope change-out. Trends of average hoist rope life and average drag rope life determined from the field study are shown in [Fig. 1] as a function of bucket size. The three lines on the graph have slopes of 1, M, and 1/4. A comparison of the data shows that hoist ropes moved about twice as much material as drag ropes prior to retirement, although more than a two-to-one scatter within each category is evident. As expected, the larger machines with the larger buckets moved more total volume of material. The four pairs of solid symbols represent those data for which detailed rope life histories are presented in subsequent figures.

Jan 1, 1980

By William R. Henkle

High-calcium limestone has been intermittently produced from the Monarch quarry since the late 1800s. Initial production was used in sugar processing and for a smelter located near Salida, Colorado. During the 1920s, Colorado Fuel and Iron Company (CF&I) began to exploit the deposit as a source of flux stone for its blast furnace steel operations at Pueblo. The deposit was operated for many years as a captive quarry by CF&I. In 1984, CF&I closed the quarrying operations at Monarch; flux stone was no longer required since the blast furnaces were taken off line in 198 1. In 1994, Colorado Lime Company, a subsidiary of United States Lime and Minerals, Inc. acquired the property. Limestone was mined from the Castle Butte Member of the Leadville Limestone (Lower Missippian). Some high- magnesium dolomite was also mined during WWII from the Redcliff Dolomite Member which directly underlies the Castle Butte. Past production was from both an upper and lower quarry at elevations ranging from 10,000 to 11,000 feet (3,048 m-3,353 m) mean sea level (m.s.1.). The reserves at Monarch quarry are in a dip slope deposit, occupying the northwest slope of Monarch Hill. The structural geology of the deposit is complex on both a region- al and mine scale. The reserves are located on the eastern limb of the Syncline Hill syncline. The beds on this limb of the structure strike to the northeast and dip to the northwest from 30 to 45 degrees. A major reverse fault separates the reserves into an upper and lower deposit. Numerous smaller normal and reverse faults cut the deposit, requiring a flexible mine plan. At the Monarch quarry, the Castle Butte Member con- sists of two stratigraphic units. The Upper unit (Mlu) is a thick-bedded, fine-grained crystalline dolomitic limestone from 0 to 300 feet (0-90 m) thick. The unit generally has an Si02 content of less than 4 percent. A bedded chert horizon and a karst horizon at the base of the unit locally raise the Si02 content as high as 20 percent. The underlying Quarry Stone unit (Mlq) is a light-gray, massive, bedded, calcilutite/micrite limestone from 90 to 130 feet (27-40 m) thick. The Quarry Stone unit is remarkably uniform chemically throughout it's entire thickness. Average analysis is CaC03 - 95 percent, MgC03 - 3.5 percent, and Si02 - 0.5 percent. The Footwall dolomite unit (Mld) of the Redcliff Member of the Leadville Limestone is a dark gray to black, blocky, medium-to-thick-bedded, calcilutite/micrite dolomite from 150 to 180 feet thick (46-55 m). The upper 100 feet (30 m) of the unit is a low-silica facies with an average Si02 content of less than 0.5 percent. MgC03 content is con- sistent at 39.8 percent throughout the upper 50 feet (15 m) of the unit but varies from 15 percent to 39.8 percent in the lower portion (50 to 100 feet [15-30 m]) of the low-silica facies. The lowest portion of the unit, 100 feet (30 m) to base of unit is high in silica and is not considered chemical grade. Current in-place, surface-minable reserves at the quarry are estimated to be 14 million short tons (st) (13 million metric tons [mt]) of limestone and 50 million st (45 million mt) of dolomite. A 1.4 million st (1.3 million mt) stockpile of finely crushed limestone is also present on the property. Colorado Lime is presently processing approximately 2,000 st (1,800 mt) per month from the crushed limestone stockpile. This material is pulverized and sold to underground coal mines for use as rock dust, some of this product is also used for environmental purposes. Limestone mining is scheduled to resume in spring of 2000.

Jan 1, 2001

Kidd Mine in Northern Ontario is now producing from Stage 1 of the Mine D expansion project ? mining between 2100 m and 2700 m below surface. Planning for Stage 2 of the Mine D project is already well advanced and will see the mine extended potentially to 3100 m below the surface. To ensure adequate ventilation is maintained at such depths, two large axial fans have been installed on the 6000 Level (1800 m below surface) as exhaust booster fans to operate in series with the surface exhaust fans. The Kidd Mine 6000 Level Booster Fans are a unique installation with unique and often unexpected challenges. These challenges have been met and attacked through the combined efforts of mine and fan manufacturer personnel with a common focus.

Jan 1, 2009

By Bruce Richardson

A contractor-manufactured 4.75m (15'-7") diameter EPB shield excavated a 115m (375 ft) long tunnel under the San Antonio River as part of the Salado Creek to Dos Rios Pipeline Interconnect Project in San Antonio, Texas. Lignite, limestone concretions, and sandstone in the face of the tunnel required the use of disc cutters. Less than 7.6m (25 ft) of sand and gravel hydraulically connected the top of the tunnel to the channel of the river. Fully lagged steel ribs with filter fabric were used for initial ground support and water control. This and other contractor manufactured TBMs in the 1.2 to 4.75m (4 to 16 ft) range are presented with descriptions of the jobsites, equipment, performances, and problems.

Jan 1, 2008

By Frederic Chavan, Hans Jakob Windier, Shane Yanagisawa, Willy Ritz

The Zurichberg twin track tunnel is the main segment of the Swiss Federal Railway's new expressline. A heavy concrete lining with a circular cross-section was chosen to cope with the swelling tendency of the surrounding rock. The 4,356 m (14,300 ft) long, 11.52 m (38 ft) diameter tunnel was driven using a shielded Robbins TBM. A 1.20 m (4 ft) wide, 25 cm (10 in) thick precast lining was placed from within the shield tail as the excavation progressed. Tunnel excavation and initial lining operation were completed in 18 months, averaging 11.86 m/day (40 ft/day) in two nine hour shifts. After breaking through, the final 28 cm (11 in) thick concrete arch lining was cast in place.

Jan 1, 1989

By Jon Y. Kaneshiro, Stephen J. Navin, Jerry Ostberg

In 1968, the U.S. Bureau of Reclamation awarded S.S. Mullen Construction of Seattle, the Intake Tunnel (488 m by 4 m diameter tunnel) and Pumping Plant No. 1. In1998, the Southern Nevada Water Authority awarded Lake Mead Constructors (Kiewit of Omaha) a parallel New Intake Tunnel (518 m long by 4.6 m diameter tunnel) and Pumping Plant. Both projects provide water to a thirsty Las Vegas. This paper presents a detailed comparison of the two intake tunnels in terms of grout takes and pressures, geologic mapping and rock quality, grout penetration beyond the excavation line, degree of fracture grouting, and groundwater inflow boundary considerations. The paper provides a comparison of grouting techniques and material developments then and now.

Jan 1, 2001

By H. Vigil Fernandez, G. Peach

The Musaimeer pump station and outfall project located in Doha Qatar, was designed to collect and manage ground and stormwater and then discharge the treated waters offshore. The project constructed a long outfall tunnel using an EPB TBM with segmental lining, each tunnel ring being 1.35 m in length with an internal diameter of 3.7 m.. The diffuser structure was a 6-arm structure, measuring 294m by 40 m connected to a central manifold in turn connected to the TBM constructed tunnel via a riser shaft drilled from an off-shore barge. This technical paper will discuss the planning, methodology and construction practices employed to connect the TBM tunnel to the diffuser structure 10.2 km offshore.

Jun 13, 2023

By W. S. Versluis, G. W. R. Kuzyk

Atomic Energy of Canada, Ltd. (AECL) is constructing an Underground Research Laboratory (URL), a major component of-the Canadian Nuclear Fuel Waste Management Program. In July 1988. AECL and the US Department of Energy commenced a cooperative research program that included extending the URL shaft by 190 m (623 ft) to the 443-m (1,453 ft) depth. The shaft extension, 4.6-m- diam (15-ft-diam), was excavated by full-face blasting using a burn cut. A Galloway stage accommodated the research activities and the sinking contractor's requirements. Blast designs incorporated controlled blasting techniques to minimize blast-induced damage on the shaft walls. The full-face method has been very effective, with 3.5-m (11.5-ft) rounds being advanced.

Jan 1, 1990

By Stephen H. E. Phillips, Michael J. Beus

A pneumatic impact breaker is used to fragment the rock at the shaft bottom, replacing the drill and blast cycle. A backhoe-type mucker removes the broken rock. The operators are in protective cabs below the bottom work deck in the shaft. Preliminary testing has been accomplished at underground and surface test sites in Arizona. Seven, approximately 6.4 m deep by 7.3 m-diameter (21 ft by 24 ft) test holes, (a total of 45 m (148 ft)), were recently excavated in various rock types, including jointed and massive quartz monzonite and rhyolite. An average advance rate of .4 m/hr (1.24 ft/hr) has been achieved. This includes breaking and mucking the rock. Further data is being sought from an in-shaft demonstration.

Jan 1, 1981

Two 6.17-m-diam (20.25-ft) shafts, said to be the largest ever opened with mechanical raise boring techniques, were completed in May and June 1978, at Monterey Coal Co.'s No. 1 mine near Carlinville, Ill. According to the shaft contractor for the project, Frontier-Kemper Constructors, the openings are 60% greater in cross section than the largest previous raise-bored shaft, which had a diameter of 4.85 m (16 ft). Both vertical shafts are 93 m (305 ft) in total depth, with the upper 28-30 m (90-100 ft) sunk and lined by conventional methods through unconsolidated overburden and the remainder mechanically bored through sedimentary rock. The two shafts will be used for service access as well as to augment ventilation.

Jan 12, 1978

By M Uceda, E Guerra, Y Choi, J L. Shepherd

In this paper we describe the leaching characteristics of pure gold in thiosulfate solutions that contain a known chelator; meso-2,3-dimercaptosuccinic acid (DMSA). The interfacial and chemical characterisation of the system was accomplished using both electrochemical methods and inductively coupled plasma mass spectrometry (ICP-MS) analysis after pure gold samples were exposed to solutions composed of 65 mM Na2S2O3, 1 mM NaOH and a range of DMSA concentrations. In the absence of DMSA, cyclic voltammetry revealed the onset of thiosulfate mediated gold dissolution, the degree of which was reduced at more positive potentials due to the formation of oxidised thiosulfate products on the gold surface. The addition of small amounts of DMSA (<10 ¦M) to the electrolyte resulted in minor changes to the voltammograms indicating a subtle interaction of the additive with the gold surface that may improve leaching. However, at higher concentrations of DMSA (up to 100 ¦M), electrochemical evidence suggests the formation of new oxidation products in solution as well as the formation of more compact monolayers of DMSA on the surface, which contribute to partial repassivation of the surface. The influence of DMSA on thiosulfate leaching of pure gold electrodes at the open circuit potential for a period of 90 minutes was tested and the corresponding gold leaching performance evaluated with ICP-MS analysis. Our data shows that small amounts of DMSA (1 ¦M) doubled the amount of total gold in solution when compared to the same study in the absence of the additive. However, higher concentrations (5 ¦M DMSA) reduced the overall amount of dissolved gold. These results may point to a mechanism where small amounts of DMSA improve leaching by adsorbing onto the surface in a highly defective monolayer that partially blocks or disrupts the formation of passivation products caused by thiosulfate oxidation, but excess DMSA can limit the leaching process itself by forming too coherent a monolayer on the gold surface.CITATION:Uceda, M, Choi, Y, Shepherd, J L and Guerra, E, 2013. Investigating the influence of meso-2,3-dimercaptosuccinic acid on thiosulfate leaching of bulk gold electrodes, in Proceedings World Gold 2013 , pp 207-214 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 26, 2013

By Gilles Cachia, Foteini Vasilikou, Seth Pollak, Hannes Lagger

"INTRODUCTION The Shatin to Central Link (SCL) is a new rail line being constructed by the MTR Corporation for the Government of Hong Kong to cater for the growing passenger demand between the North-East New Territories and the urban areas of the city. Contract 1103, which will provide the 4 km of running tunnels between the new Hin Keng and Diamond Hill Stations, was tendered in a design-build framework and awarded to Vinci Construction Grands Projets (VCGP) in October 2012, with Arup as Vinci’s lead designer. One of the most challenging pieces of work on the contract involved the construction of 133m of SEM mined tunnel on the northern end of the project. The Northern Mined Section (NMS) is an undrained, twin-track, single bore profile, broken out through the southern pipe pile wall of the Hin Keng Station cut-and-cover box, Figure 1. The tunnel has maximum excavated dimensions of 14.5m span x 12m height, with track spacing of 6.175m. The general excavation sequence consisted of 3 stages; a 6m high top heading (73.1m2), followed by bench (27.3m2), and invert (26.7m2). A risk-based approach to the design was taken given the large variation in rock level, cover, and an active stream course above tunnel crown. Ultimately, six different temporary support types were utilized in order to optimize the construction process through continually changing ground while maintaining a safe and stable excavation. Three main support types were developed during the initial design phase, along with a “contingency” case, which utilized a partial heading excavation sequence and enhanced face support. The purpose of the contingency support type was to have a robust, engineered design available for immediate implementation should unexpected ground behavior develop, leading to the breaching of instrumentation trigger levels. This support type was never implemented on site. Drill- and-blast tunneling commenced once 6m of rock cover was verified through forward probing. A robust instrumentation and monitoring scheme provided valuable data throughout excavation and the means to carry out a geotechnical back-analysis of the ground behavior and support performance."

Jan 1, 2016

By Lance Rasband, Richard Hanke, John Starcevich

"Mining on Seattle’s SR99 tunnel began in July of 2013. Initial mining progressed slowly through a 4.5 m (14.9-ft) thick fiber-reinforced concrete headwall before immediately entering the first of three planned safe havens. Once through the headwall and safe haven the machine would move through roughly 122 lineal m (400 ft) of jet grout improved soil at a rate of slightly more than 1.6 m/d (5.3 ftpd). It was anticipated that once mining has passed through the improved area, “Bertha” – now the world’s second largest tunnel boring machine (TBM), would be moving at a rate of almost 11 m/d (36 ftpd). By the end of October 2013, Bertha had mined roughly 131 m (430 ft), averaging 1.4 m/d (4.7 ftpd), and had begun mining within native soils. By the start of December 2013, tunneling had reached the 244 m (800 ft) mark and Bertha was averaging 3.6 m/d (11.9 ftpd), achieving as much as 12.8 m (42 ft) of progress on given days. Unfortunately, on Dec. 6th, 2013, Bertha began to stumble and all progress quickly came to a halt.Initial reports stated an obstruction had stopped Bertha in her tracks. The preliminary investigation involved installation of 10 deep dewatering wells to reduce the hydrostatic pressure within the ground around the TBM, to allow the contractor’s personnel to inspect the cutter head from within and also to investigate for the presence of obstructions at the face of the cutter head. During inspection, the contractor’s crews found fragments of steel pipe entangled in the cutter head. The steel was apparently from an abandoned 254 mm (0.9 in.) diameter monitoring well casing, previously installed by the Washington State Department of Transportation (WSDOT) to monitor groundwater conditions for the design and planning of this tunnel project. After the steel was removed from the cutter head and cleared from Bertha’s path, she attempted to move forward but was only able to generate a slow crawl."

Jun 1, 2016

By Junji Shibata

Waste solution from the Ni non-electric plating process contains 4.0 kg/m3 Ni, 0.1 kglm3 Fe, 0.0 I kg/m3 Zn, 48 kg/m3 S042-, 98 kg/m3 HP032-and 31 kg/m3 laetic acid as a typical composition. Solvent extraction, cementation, ion exchange resin and precipitation methods are possible to use for the treatment of this kind of solution. In this study, solvent extraction of Ni was investigated using two kinds of extractants to clarify the relation between extraction percent and pH for various extractant mixtures. As Ni ion has an octahedral structure with ligands, the synergistic effect for Ni extraction is considered to take place in solvent extraction using two types of extractants. The investigation of the synergistic effect for Ni extraction was carried out in case of various extractant systems.

Jan 1, 2003

Agnico-Eagle Mines Ltd. comprises a Gold Division at Joutel, Quebec, and a Silver Division at Cobalt, Ontario. In 1979 the Silver Division milled 41,700 mt (46,000 st) of ore at a recovered grade of 139 g per mt (4.06 oz per st) for a production of 5,816 kg (187,000 oz) of silver. The 100 -year -old mining, camp of Cobalt has long been noted for its high-grade silver ores, and now seems to be on the threshold of being revived with several companies operating in the area. In 1979, the Gold Division milled at the rate of 27,200 mt per month (30,000 st) and processed 333,000 mt (367,000 st) at a grade of 6.58 g per mt (0.192 oz per st) with a mill recovery of 91.670 to produce 2,013 kg (64,700 oz) of gold. The average total cost per ton milled was $22.10 per mt ($20.05 per st) which was $3.66 per g ($1 13.90 per 0s). These figures make Agnico-Eagle one of the lowest cost producers in North America. Gold occurrence in the Joutel ore is fine-grained and contained within sulfide facies composed mainly of pyrite with minor pyrrhotite. Two underground mining methods are employed with long hole stoping accounting for about 75% of the ore mined and shrinkage stoping the remainder. Ore from underground is hoisted into a 545 mt (600 st) coarse ore bin from where it is conveyed to a 45 mt (50 st) surge bin ahead of the 10-51 Hydrocone crusher. Crushed ore is then screened to make a 13 mm (112 in.) product with the oversize being recrushed in a 7-60 Hydrocone. Mill feed is conveyed to two 900 mt (1,000 st) fine ore bins. The Hydrocones were installed as part of a new crushing plant in March of 1980 and have contributed substantially to an increase in mill capacity. Space has been left to install a third crusher in the future. The fine ore bins discharge to a 2.75 m by 3.66 m (9 ft by 12 ft) rod mill, followed by a 3. 7 m by 5.5 m (12 ft by 18 ft) ball mill in closed circuit with a Krebs 380 mm (15 in.) cyclone to produce a 7070 minus 37 micron (400 mesh) .flotation feed. This pulp is first treated in a 34 cu m (1, 200 cu ft) Maxwell flotation cell and then in an additional twelve 2.8 cu m (100 cu ft) cells. Combined flotation concentrates are reground to 9070 minus 37 microns (400 mesh) in a 1.8 m by 4.9 m (6 ft by 16 ft) ball mill in closed circuit with a single Krebs 254 mrn (10 in.) cyclone. Flotation tailings are sent to the tailings dam. Reground concentrates are then aerated with lime to a single Pachuca and thickened in a 16.75 m (55 ft) diameter thickener with the addition of diesel fuel to render any activated graphite inert to adsorption of

Jan 1, 1981

By Kazutoshi Haga, Batnasan Altansukh, Atsushi Shibayama

In this study, the dissolution behavior of copper and arsenic from five different copper concentrates containing 21-34.7 %wt. of Cu, 1.3-7.3 %wt. of As and 15.6-27.5 %wt. of Fe were investigated under high pressure oxidative leaching conditions. Leaching experiments were conducted using different leaching reagents such as ferric sulfate, sodium chloride, deionized water, sulfuric acid and ferric sulfate-sodium chloride mixture in order to compare the effectiveness of each on the dissolution of copper and arsenic from copper concentrates. The effects of ferric sulfate (0.1-1 mol/L), sodium chloride (0.1-0.5 mol/L) and leaching temperature (100-180C) were investigated under 1 MPa autoclave pressure , particle size >45 m, pulp density of 100 g/L, stirring speed of 750 rpm and leaching time of 60 min. Results showed that the most effective leaching reagents for the dissolution of copper are 0.34 M NaCl and a mixture of 0.1 M Fe2(SO4)3 - 0.34 M NaCl, in which As extraction did not exceed 7%. The copper concentration in the pregnant leach solution obtained after leaching using 0.1 M Fe2(SO4)3 - 0.34 M NaCl was 13-30 g/L, while the concentration of As and Fe were in a range between 0.3-1 g/L and 0.2-10 g/L at 160C. INTRODUCTION Copper sulfide minerals, such as chalcopyrite (CuFeS2), chalcocite (Cu2S) and covellite (CuS) are used for the production of copper in mining and metallurgical industries worldwide. The conventional pyrometallurgical processing followed by electrorefining is a common process for the production of high purity copper (Moskalyk, & Alfantazi, 2003; Padilla, Aracena, & Ruiz, 2012). However, the degradation of copper content and an increase in the amount of metal impurities, especially arsenic (As), antimony (Sb) and bismuth (Bi) in copper concentrates cause serious problems to the production of copper and environmental pollution (Basha, Selvi, Ramasamy, & Chellammal, 2008; Long, Peng, & Bradshaw, 2012). Hydrometallurgical processes are used to recover metals from refractory ores, concentrates and tailings by leaching (Long, Peng, & Bradshaw, 2012; Masloboev, Seleznev, Svetlov, & Makarov, 2018). During the last decades, hydrometallurgical processing for the recovery of copper from enargite-bearing copper ores and concentrates has been extensively investigated due to its relatively more environmentally benign and less capital costs than the conventional pyrometallurgical processes (Long, Peng, & Bradshaw, 2012; Masloboev, Seleznev, Svetlov, & Makarov, 2018; Choubey, Lee, Kim, & Kim 2018). However, several issues facing the hydrometallurgical processing are incomplete extraction of copper associated with the coprecipitation of iron compounds, difficult stabilization of toxic impurities and production of a contaminated copper product (Dreisinger, 2003; Safarzadeh, Moats, & Miller, 2014). Hence, a number of studies have been conducted on the selective dissolution of copper from copper ores and concentrates or selective removal of arsenic leaving a clean copper residue using various leaching reagents or lixiviants (Safarzadeh, Moats, & Miller, 2014; Mihajlovic, Strbac, Zivkovic, Kovacevic, & Stehernic, 2007; Tongamp, Takasaki, & Shibayama, 2009). More extensive studies on the selective extraction of copper from enargite- bearing copper ores and concentrates are much needed to improve copper recovery.

Jan 1, 2019

By B. Simser

At Brunswick Mine, an increase in the number of seismically-related falls of ground and in particular rockbursts occurred during 1999. This was due to the ever-increasing stress levels in the remaining ore reserves, mining in difficult highly stressed areas and in particular troublesome geological features. In conjunction with the Noranda Technology Centre, development of rockburst support systems started at the Brunswick Mine in 1996 and efforts were accelerated in 1999. A support package consisting of #6 gauge chain link mesh, #0 gauge heavy mesh straps, and a 1 m by 1 m pattern of modified conebolts was developed. The conebolts were modified to enable mixing of the resin anchoring. Two major rockbursts occurred during an intense seismic episode at the Brunswick Mine from October 13th to 17th , 2000.The area was evacuated approximately four hours before the first burst hit, and remained closed during the seismic ?flurry??. The first damaging activity occurred around midnight on Friday the 13th, and a second damaging episode occurred on the 17th that included a local Richter Magnitude 2.7 event. The second rockburst had dimensions of approximately 5 m by 5 m by 20 m or 2150 tonnes of massive sulphide material (density 4300 kg/m3). The first failure area was more extensive, caving approximately six metres of an intersection back with the lateral extent not safely measurable. This paper briefly describes the mining conditions that led to the bursting, the seismic response of the region, the mechanisms thought to have caused the bursts, and a yielding support system which had been partially installed before the burst.

Jan 1, 2002