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By P A. Tilyard

There have been enormous changes in mineral processing in the past four decades. For example grinding mill power has increased by an order of magnitude, regrinding is done to -10 microns and flotation machines are 100 times bigger. Operating staff have unprecedented opportunities for online monitoring and performance control of mineral processing plants. Sophisticated instruments can provide a plethora of data characterising the mineralogy and surfaces of particles. Digital computers allow complex calculations on huge amounts of data including modelling and simulation of machine and plant performance. However, all these changes have not necessarily lead to better metallurgical results. An analogy can be drawn with the thoroughbred racing industry in Australia. Significant advances in scientific knowledge in animal genetics, physiology, biomechanics and nutrition applied to the business have resulted in only a two per cent reduction in winning times for the Melbourne Cup and Caulfield Cup since the 1920s. A critical look at some mineral processing metrics suggests similar failures to improve performance despite putting in more resources. In fact certain parameters such as operating times and plant start-up performance are considered to have remained static or even deteriorated. There has been an emphasis on 'process' at the expense of 'outcomes'. The industry's strength has been in finding technical (or 'hardware') solutions while its weakness has been at the people end of the business in maximising and consolidating the gains from the technologies. Some trends in plant design over these years have exacerbated the apparent deskilling of operating and technical staff. Despite unparalleled options for communications, some staff are embarrassingly uninformed about technical developments in their fields. The 'boom and bust' cycles of the industry, together with trends in tertiary education and the effects of fly-in, fly-out (FIFO) operations, raise serious questions about the sustainability of human capital in the mineral processing sector. This paper by two experienced mineral processing engineers, with contributions from other senior practitioners, reviews these trends. While there may be an element of 'the older we are, the better we were', it is an attempt to identify the issues and propose solutions.

Jan 1, 2009

By Andreas Dietrich

The development of new techniques in mineral exploration seems to reduce the need of geological field work and mapping: Data collection can now be done remotely by satellites, drones, and tablet-equipped field assistants, while the geologist assembles data sets with 3D software. The interpretation of complex data sets can be assisted by machine learning and artificial intelligence (ML, AI). Instead of loosing relevance, geological mapping and logging are now even more essential to provide the interpretive framework essential for the correlation, validation, and integration of huge amounts of data, but will not be replaced by new technologies. Geological truthing lies in the field, with experienced geologists conducting this work while training and mentoring young professionals.

May 10, 2023

By J. Arzúa, L. R. Alejano, P. Rodríguez

"After six years of successful excavation with a support pattern, some roof failures have happened in a room-and-pillar underground mine excavated in a stratified rock mass. In this paper the authors give explanation to the failure mechanisms and redesign the support in order to solve the recent problems of roof collapse.The mine magnesite bed is approximately 14 m thick, and runs in the E-W direction with an average dip of 18º. Due to mine constraints and the inexpensiveness of the ore, the excavation is performed in the mineral itself, avoiding the need for mining dumps and waste excavation. Rooms and drifts are oriented in the mineral bed to form an angle of 19º to the direction of the strata, reducing the roadway gradients to less than 10%. Rooms are 11 m wide and pillars are 7 m wide. They are designed to leave a 1 meter thick magnesite bed in the roof of the rooms to act as a supporting beam. More often than not there is a layer of very poor quality marly shale up to 1 m thick resting on this magnesite beam. Overlying the poor quality marly shale there is a better quality, self-supporting sandy shale.First, a characterization of the involved rock masses (magnesite, marly shale and sandy shale) is performed, collecting the geotechnical data from previous studies and performing new in situ geotechnical surveys to check the soundness of the previous data. A distinction is made between different quality zones inside the mine by means of the RMR classification, associated to the occurrence of faults.Back-analysis of some of the failed rooms and data from exploration boreholes were used to determine the failure mechanisms (associated to Voussoir type failure or occurrence of sub-horizontal joints crossing the magnesite beam). Based on an understanding of these mechanisms, the method used to design the support and to analyse the stability of the different possible thicknesses of the magnesite beams of the rooms and drifts of the mine were improved.Numerical models of the most relevant conditions found in the mine were performed in order to better understand the behaviour of the roofs and to simulate the supporting solution obtained in the previous analyses.The most relevant conclusions with regard to the main elements causing the instability of the roof include the appearance of sub-horizontal discontinuities crossing the magnesite bed, which causes total instability of the roof and usually leads to collapse; excessive thinness of the magnesite beam in the problematic rooms; and the relaxation of the deepest (around 270 m in depth) zone of the mine because of existing faults."

Jan 1, 2015

By J. F. T. Agapito

Experience at Energy West Mines in the Wasatch Plateau Coal Field, Utah, has shown that longwall face bursts occur mostly near the tailgate when extracting the third adjacent longwall in a panel section where depths exceed 610 m (2000 ft). It is believed that this is due to a combination of cover depth and side abutment stresses created by the widening of a pressure arch formed by interlocking and bridging of large blocks of" very competent strata. The 60-m-thick (200 ft) Castlegate Sandstone probably plays a key role in pressure arching. This paper presents a back-analysis of the location and magnitude of the stresses associated with face bursts that occurred when mining the 7th Right Longwall at the Deer Creek Mine. Improvement of these conditions may be achieved by widening the longwalls to attain critical subsidence width sooner while mining the second long-wall, and to reduce stresses at the face. Wider longwall faces, however, increase the chance of sandstone channel interception often associated with bursts. Satisfactory mitigation of face bursts beyond cover depths of 610 to 670 m (2000 to 2200 ft) may be difficult without using destressing techniques. The hydraulic fracturing/liquid infusion technique seems to offer the best possibilities for application in longwall mining.

Jan 1, 1997

By G. Russo

During the construction of the S.Martin La Porte access adit, 9m in diameter, extreme squeezing conditions occurred while crossing Carboniferous schistose formations, with diametric convergence up to 2m. Back-analysis of monitoring results shows that an acceptable simulation of the observed excavation behaviour can be derived using a combined analytical and numerical approach, with simple elasto-plastic laws but a progressive reduction of residual parameters and a relevant increase of dilatancy.

May 1, 2009

By E. Moriguchi, K. Tamaki, A. Okuda, K. Iizasa, C. Honsho

The northern Izu-Ogasawara Arc is characterized by an abundant distribution of submarine calderas along the volcanic front to the back-arc rift zones. In many of these calderas hydrothermal activities are observed. Among these calderas, hydrothermal ore deposits with economically feasible quality and amounts are hosted at two calderas, the Myojin Knoll Caldera and Beyonnaise Knoll Caldera. We conducted two research cruises in 2005 and 2006 to investigate further hydrothermal activities and ore deposits in these calderas and to complete the geological and geophysical mapping in the area. We discovered a strong signature of hydrothermal activity in the Myojin-sho Caldera using AUV capabilities. Newly obtained geological and geophysical data clarified the relationship between back-arc rifting tectonics and hydrothermal activities in the area. Five calderas (two of them newly found) and numerous (>200) volcanoes are mapped in the area. Complete SeaBeam coverage showed a distinct feature of the back-arc rifting activity and its relationship to the submarine calderas on the volcanic front. The back-arc rift zones composed of four segments separated by E-W trending volcanic ridges. Three of the calderas are located in the rift zone and are heavily deformed by the rift normal faulting that typically trend NNW-SSE. As the Beyonnaise Caldera is located in a swarm of faulting zones at the eastern end of the back-arc rift, the formation of the hydrothermal ore deposits found in the Beyonnaise Caldera are apparently affected by these faulting activities. Newly found submarine calderas are located in the rifting zone and are heavily faulted.

Sep 24, 2006

By Torsten Wichtmann, Jan Macheck, Frederik Broz, Patrick Staubach, Hauke Zachert

The majority of heavy or strongly loaded structures, for instance high-rise buildings, large span bridges or gas storage tanks are founded on pile groups. However, current design codes and guidelines do not provide a standard procedure for their design under lateral dynamic and high-cyclic loading arising from e.g. wind, temperature or earthquake loads. To establish appropriate design strategies, a better understanding of the pile-soil, pile-pile and pile-cap interaction is required. This paper is focused on finite element backcalculations of centrifuge tests on single piles and pile groups subjected to up to 500 lateral load cycles. For the finite element simulations, a special calculation procedure is applied, combining two kinds of constitutive models, a conventional (Hypoplasticity with Intergranular Strain extension) and a special highcycle accumulation (HCA) model. The parameters of the constitutive models are calibrated based on static and cyclic laboratory tests on the sand used in the centrifuge tests. The performance of the applied numerical procedure is evaluated based on a detailed comparison of the simulation results with the measured data from the centrifuge tests. Overall the applicability of the used numerical strategy for simulating the behavior of pile groups subjected to high-cyclic loading is shown.

May 1, 2022

By Howard Talbot

THE history of the development of the strip in- dustry, mentioned in this paper, was covered in considerable detail by Stephen Badlam in his paper presented. before the American Iron and Steel Institute at its meeting in New York last October. The description of progress in rolling strip there out-lined will be found good reading by anyone interested in the subject. Quoting from Mr. Badlam's paper, "For practically thirty years, we find no record of the use of the backing-roll principle, until in 1916 the Lukens Steel Co., Coatesville, Pa., put in a 206-in. plate mill of the four-high type. This was a reversing mill with two 32-in. working rolls and two 50-in. backing rolls and is today making the widest and heaviest plate made anywhere in the world." The functioning of this mill, though rolling hot plate, is, on account of the extreme width and thinness of plate rolled, quite comparable to the rolling of wide strip either hot or cold and was a considerable factor in demonstrating the marked advantages available with the four-high or backing-roll mill for thin accurate rolling. As the writer has brought out, a demand for a product, the rolling of which required the character-istics of the backing-roll mill, was broached by the In-ternational Nickel Co. This resulted in an inquiry of March, 1924, for the building of a mill capable of cold-rolling wide nickel and monel metal sheets or strip, as mentioned. It may be of some interest to note, that whereas the .six-roll cluster mill with plain bearings was built for this installation, the mill designed by the United En-gineering and Foundry Co. for this inquiry was of the four-high type with roller bearings on the backing-roll necks, being essentially the same design as the four-high roller-bearing mill built in 1925 for the Rome Brass & Copper Co. This Rome mill, to quote again from Mr. Badlam's paper, "fulfilled every requirement, and was the model on which succeeding mills of this type were built."

Jan 5, 1928

By Lloyd Jones

THE strip industry made rapid strides in regard to both width and gage until about 1922, when the maximum width was about 20 in. In the hot mills, strips of thin gages in wide widths could be pro-duced, but only at a sacrifice of length of coil, con-siderable increase in power, and decreased tonnage out-put on the min. Hot mills, like the Bray sheet mill and the modern strip and skelp mills, demonstrated that . wider strip could readily be rolled, if such a product was desired. The strip manufacturers were really limited in width, not by the hot strip, but by the cold rolling of wide strip. This factor was roll pressure. The diameter of rolls increased as the width of strip increased. For cold rolling strip 20 in. wide, cold rolls were approximately 20 in. diameter. As the rolls became larger, the arc of contact between the rolls and the strip increased approximately, according to the formula, b=\/Rd where b = the width of contact R = roll radius d = total reduction Now the crushing strength of cold steel of low carbon varies from 70,000 to 100,000 lb. per sq. in. With each succeeding pass the crushing strength increases and may approach 250,000 lb., according to the total reduction. The total reduction in cold rolling of low-carbon steel will range from 40 to 70 per cent of the original thickness, usually in about three or four passes. On this basis, assuming a reduction of 20 per cent in one pass, the crushing strength of steel assumed at 120,000 lb., we have a total pressure between rolls, which (assuming that the rolls do not deform) equals: Roll Strip Strip Diameter, Width, Thickness, In. In. In. Pounds 10 10 0.100 380,000 20 20 0.100 1,080,000 30 36 0.100 2,375,000 42 50 0.100 3,900,000 These figures are approximate, but you will note that in increasing the width five times, or 50 in., the total pressure between the rolls increased approximately ten times. In other words, the pressure per inch of width increased from 38,000 to 78,000 lb. for the same per-centage of reduction.

Jan 3, 1928

By Dan Kump

This paper provides an overview of Barrick Gold's Bullfrog underground mine backfill program. The Bullfrog underground mine was a gold mine that primarily used the underhand cut-and-fill mining method. The first portal was established June 18, 1991, and mining continued until Dec. 29, 1998. This paper begins with a discussion of the events that took place to reach about 1.8 kt/d (2,000 stpd) of backfill. It ends with a commentary of what was learned. The Bullfrog experience suggests that there are two essential factors to a successful backfill program: a well-graded aggregate supply and proper placing procedures.

Jan 1, 2002

By D. Kump

The Bullfrog Mine is an underground gold mine that uses the underhand-cut-and-fill mining method. The portal was established on June 18, 1991. Mining is expected to be completed by the summer of 1999. This paper starts by discussing what it took to get from nothing to 2000 tons-per-day backfill, and ends with a hindsight commentary on things that did and did not help.

Jan 1, 1999

By G York

South African gold reserves extend to great depths. However, strategies for the mining of such deposits have changed only slightly over the past number of years. Regional support at depth is currently provided by rock pillars and backfill. Many problems have been experienced with these regional support practices at deeper levels. In this paper several alternative forms of regional support are described which include the use of concrete, the use of back area caving and confinement to strengthen pillars. These new concepts are supported by numerical and laboratory modelling and in situ results.

Jan 1, 1998

By L. H. Van Loon, E. G. Crayston, A. F. Heather

"IntroductionTHE PROPERTY of Madsen Red Lake Gold Mines. Limited. comprises fifty-eight claims located between Russet and Faulkenham lakes in Baird and Heyson townships of the Patricia portion of the District of Kenora. The mine is seven miles •by road from the hamlet of Red Lake. This latter centre is .connected to the Trans-Canada Highway at Vermillion Bay, Ontario, by provincial highway No. 105 which has a total length of one hundred and thirteen miles.The Madsen property went into production on August 11th, 1938, with a milling plant rated at 300 tons per twenty-four hours. Changes made in the crushing plant during October, 1938, allowed a •finer feed to be delivered to the mill and increased the milling •capacity to 400 tons per twenty-four hours. During 1948, steps were taken to add an additional 400-ton capacity unit to the milling plant. This latter work was completed by May, 1949, giving the present total milling .capacity of 800 tons per twenty-four hours.It was realized in the early 1940's that, eventually, considerable material suitable for backfilling in the mine would be required. As there were no large deposits of gravel or sand anywhere near the mine, officials of the Company considered it advisable to investigate the feasibility of using mill tailings for mine •backfill. Since Madsen mill tailings contain approximately 80 per •cent of minus-200-mesh material, it was not considered practical to use them for underground backfill without first removing the finer material. as the slimes present in a product of this fineness would without doubt preclude any drainage whatsoever. Table I gives a sizing analysis of Madsen mill tailings and Table II a quantitative chemical analysis of the ore."

Jan 1, 1954

By Graham Swan, Richard K. Brummer

The paper begins by describing an underhand cut-and-fill mining concept for a 50m thick, generally tabular though irregularly shaped orebody dipping at 30 degrees and occurring at a depth of between 2300 and 2700m in the Sudbury Basin. From this, the broad design requirements of a fill system are stated. Two and three-dimensional numerical models are then described which were built to evaluate the theoretical benefits of different types of fill reinforcement. Included in this evaluation are the effects of span, fill strength, fill tightness, as well as static and dynamic loading types. Results, supported by large-scale physical models, are presented to show the significance of reinforcement on stable, undermined fill performance. Finally the paper describes work which has been done to evaluate the suitability of so-called silicated cemented backfill. This system is, in principal, attractive for deep mining applications since transportation is in the "safe" hydraulic regime, while fill water run-off, segregation and shrinkage effects are potentially eliminated in the stopes. The results from various surface and underground trial pours are reported which substantiate the proposed use of this type of backfill at depth.

Jan 1, 2001

By C. A. Rawlins

In an underground mine the ingress of heat can generally be attributed to two sources (Rock exposure and artificial mechanisms). The heat load relating rock exposure is subdivided into stoping [~40%] and haulages [~20%]. Rock is exposed to the ambient air and as a mine increases in depth the VRT (Virgin Rock Temperature) increases which leads to more heat ingress underground. To counter the heat ingress, especially in low-seam stoping areas of a deep level mine, air is, a) cooled artificially, b) the amount of rock exposed is reduced in the back areas of a stope by backfill applications, and c) ventilation air is used more efficiently. The heat reduction within a stope due to placed backfill could amount to ~50% for operations mining at a mean rock breaking depth of about 4500 mbc (metres below collar). Unfortunately, due to backfill transport methods, the physical fluid properties and application methods adopted, additional heat is induced into the air. The heat ingress due to backfill application could amount to as much as 25% of the overall in-stope heat load. This heat influx, seen as a negative aspect of backfill operations, is nevertheless overshadowed by the overall in-mine heat load reduction capability of about 20% attainable. The in-stope benefit [~50%] attainable from backfill operations could be increased with the incorporation of a heat exchange (cooling) methodology to offset the additional heat ingress into the surrounding air of a deep underground mine.

Jan 1, 2003

Stope width was controlled around 3.5 feet for the narrow high grade vein system at Hollister Gold Mine. Without top access, backfill material was pumped from bottom access up into stope through a series of pipes after ore had been slushed out in each mining cycle. Backfill material was changed from one layer of low strength concrete plus one layer of high strength concrete cap to one layer of low strength cellular concrete plus one layer of high strength concrete cap. Backfill cost was cut down to 30 percent during this process. To reduce the steel concrete pipe set up and tear down time which was about 2 to 3 shifts in each mining cycle, steel pipes were left in man-way between mining cycles. Cellular concrete was firstly introduced into backfilling stopes at Hollister mine in Nevada Mining. This process including concrete pump selection, pumping pipe selection, material mixing, material mechanics and quality control were fully covered in this paper. CUT AND FILL MINING CYCLE The most flexible mining method, cut and fill (Figure 1), was used to mine the high grade, narrow steep vein system on this Northern Nevada property which is located eighty miles north east of Winnemucca. Without top access, stopes were accessed through two man-ways on each end, in which people travel ladders and material transport slides were set up (Figure 2). Each mining cycle was called a cut, from 6 to 8 feet high. Overhand sequence was used. Mining cycle started with drill and blast the middle portion of the stope, then expanded towards both ends. Broken ore was slushed into mill hole located at middle of the stope and then mucked out through bottom access into main haulage. Stope length was about 300 feet, which was restricted by slushing distance. Timber man-ways and mill hole would be raised and established for backfill and ready for the next mining cycle.

Jan 1, 2011

By Jamel Sgaoula

During 1998 COGEMA Resources Inc. announced that operations at the Cluff Lake mine will be suspended indefinitely by the end of year 2000 as a result of deteriorating uranium market conditions. To remain profitable, Cluff Lake mine started an aggressive cost reduction program. A cost reduction team was formed to look for ways to lower production costs. The focus of this team was to screen all areas and to come up with ideas to improve the economy of the mine and ensure the viability of the operations over the remaining three years of operations. Among the areas where significant savings were achieved are: (1) replacement of portion of cement with fly ash in backfill; (2) backfill optimization; and (3) new shotcrete plant installation on site for bulk in-house shotcrete production. This paper details the research work carried out by Cluff Lake mine on these projects and results of its application.

May 1, 2001

By T. R. Yu, D. B. Counter

"Kidd Creek Mines primarily employs sublevel blasthole stoping for underground mining operations. During pillar recovery in No.1 Mine and panel mining in No.2 Mine, the free standing fill exposures may reach a height of 100 m or more. Consequently, a competent consolidated rockfill has been required to fill the mined openings.Since 1976, a total of more than fifty stopes and pillars have been filled with over four million tonnes of backfill. Fill exposures, up to heights of 100 m, have stood up well after blasting adjacent to the backfill.The fill material consists of a graded aggregate, Portland cement and ground slag. Sand has been occasionally mixed in the fill material to serve as avoid filler. The aggregate is prepared in a fill crushing plant, transported underground through raises, and hauled to the filling stopes by conveyors in No. 1 Mine and by trucks in No.2 Mine.Two mixing methods are currently in use. No. I Mine utilizes the free fall of gravity; the fill materials are sprayed with cement slurry as the aggregate departs from the conveyor into a mixing culvert. The other is the use of a rotary mixer on a batch basis in No. 2 Mine.This paper describes the fill systems, filling operation s, and the backfill quality control practice, together with various testing results related to improving fill quality as well as reducing fill costs.IntroductionThe Kidd Creek minesite, located 27 km north o f the city centre of Timmins, Ontario, produces 4.5 million tonnes of copper, zinc, lead, and silver ores annually from its No. 1 and No.2 underground mining operations. Both mines employ a sublevel blasthole stoping method."

Jan 1, 1983

By Ike Isagon

Backfilling has become an integral process in underground mining operations, with varying cost-effective forms of application. This study is a part of an overall investigation into the backfill practices at Copper Cliff North Mine (CCNM). The mine is well-diversified in terms of backfill methodologies and practices. The mine employs different fill methods such as cemented and un-cemented rock fill, slag fill, and cemented hydraulic fill (or a combination of them) in order to fill the voids created by mining activities. The history of backfilling at CCNM has been dominated by cemented and un-cemented slag fill. Each fill method requires operating and engineering guidelines for a cost effective quality control system. The purpose of this paper is to briefly discuss the backfill methods and applications, operating guidelines, and engineering requirements at CCNM.

May 1, 2011

By D. Bronkhorst, J. Rheault

"Backfill practices at the Williams mine have evolved significantly since underground mining began in 1986. The use of conveyors as a means of transporting fill material to stopes has been discontinued in favour of load-haul-dump equipment. The method of adding binder to the fill product has been refined along with the placement technique. Slurry preparation has been fully automated and flyash is now included as a binding agent. Surface preparation of the fill product has evolved substantially with the quarry operation coming into full operation and the construction of pennanent crushing, screening, and storage facilities. Company personnel have replaced all contracted labour involved in backfill operations. Some changes have been made to achieve a higher standard of quality control while others were implemented as a matter of cost reduction. BackgroundThe Williams mine is one of three mines operating at the Hernlo gold deposit, located 40 km east of Marathon, Ontario, along the north shore of Lake Superior. Daily production at the mine is 6000 Ud yielding an annual gold production of approximately 500 000 oz. The orebody extends from surface to a depth of 1300 m dipping north at 70 degrees. The average orebody width is 25 m, but varies from 45 m on the eastern boundary to about 10 m at the western probable limit and at depth. Proven and probable reserves currently stand at 32 million tonnes grading 6 g/t.The Williams orebody has been divided into four separate mining zones: the A-zone, the West-of-Dyke zone, the B-zone, and the C-zone (Fig 1). Ore production originally was from the A-zone open pit operation while the A-zone and B-zone underground workings were being developed. The mining method employed underground is blasthole open stoping with delayed rockfill. Mining block dimensions are nominally 20 m along strike by 25 m high. Stope extraction is sequenced to achieve a relatively short pillar life and 100010 ore recovery. Pillars are extracted as soon as the adjacent primary stopes are two cycles ahead. Primary stopes are filled with consolidated rock fill after mining. This allows pillars to be blasted right against the consolidated fill, but also requires that the fill be self supporting to a height of 30 m.The Williams mine places approximately 1.4 million tonnes of rockfill and 50 000 tonnes of binding agents each year, with about three quarters of the fill placed being consolidated. The consolidated fill placed at the Williams mine contains Portland cement and flyash as binding agents."

Jan 1, 1994