Mining Techniques - Future

**79**

**Chapter 5**

Elongation

*Paweł Kamiński*

**Abstract**

adjustable guiding

**1. Introduction**

Polish Experience in Shaft

Deepening and Mining Shaft Hoist

Deepening of active mine shaft comprises number of specific an very difficult operations, because it calls for use of untypical devices securing hoist operation in the shaft, as well as special technology tailored to actual technology of the deepened shaft face. The Leon IV shaft at the Rydułtowy mine is one of the last mining shafts deepened in Polish coal mining from the surface, and then deepened and finally equipped with mine shaft hoist installation. This investment will allow for the construction of the exploitation level of 1150 m and the availability of further coal extraction up to a depth of 1200 m. It will guarantee the possibility of exploitation of over 65 million tons of coal and continuous operation of the mine until 2040. At the same time, for the first time in the Polish hard coal mining industry, a flexible guiding of the mining cage and skips was used, which in comparison with rigid guiding is a much cheaper solution and has many other advantages. The chapter presents most important problems and technical solution implemented during construction and deepening of the Leon IV Shaft at Rydułtowy Coal Mine in Poland.

**Keywords:** shaft hoist elongation, deepening mining shafts, shaft construction,

Hard coal mine Rydułtowy is one of the oldest Polish mines in Rybnik Coal District. Its predecessor named Charlotte started production in the year 1806 and it was one of the greatest mines at that time. The mine as the first was equipped with steam engine already in the year 1855 and it was connected with the rest of the country via railway line what facilitated coal sales. At the beginning of twentieth century, the mine in question passed through numerous crisis phases that resulted in the employment reductions, and in the year 1932 the mine was even closed for a period of 4 years. However, the mine developed during the Second World War because Germans needed big amounts of good-quality coal. In the period 1940– 1944, the employment was increased threefold up to 3582 workers. After the Second World War, the Charlotta mine was renamed as Rydułtowy mine, which belonged to various structures of Polish mining industry. In the year, the KWK Rydułtowy was joined with Anna mine forming mining plant named as KWK Rydułtowy-Anna. In the period 1990–1998 a new shaft Leon IV was sunked with diameter 8.5 m, which rarely occured in Polish mining industry. In this period, it was decided that the shaft depth of 1076.2 m would allow development of the new level 1050 m,

#### **Chapter 5**

## Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation

*Paweł Kamiński*

#### **Abstract**

Deepening of active mine shaft comprises number of specific an very difficult operations, because it calls for use of untypical devices securing hoist operation in the shaft, as well as special technology tailored to actual technology of the deepened shaft face. The Leon IV shaft at the Rydułtowy mine is one of the last mining shafts deepened in Polish coal mining from the surface, and then deepened and finally equipped with mine shaft hoist installation. This investment will allow for the construction of the exploitation level of 1150 m and the availability of further coal extraction up to a depth of 1200 m. It will guarantee the possibility of exploitation of over 65 million tons of coal and continuous operation of the mine until 2040. At the same time, for the first time in the Polish hard coal mining industry, a flexible guiding of the mining cage and skips was used, which in comparison with rigid guiding is a much cheaper solution and has many other advantages. The chapter presents most important problems and technical solution implemented during construction and deepening of the Leon IV Shaft at Rydułtowy Coal Mine in Poland.

**Keywords:** shaft hoist elongation, deepening mining shafts, shaft construction, adjustable guiding

#### **1. Introduction**

Hard coal mine Rydułtowy is one of the oldest Polish mines in Rybnik Coal District. Its predecessor named Charlotte started production in the year 1806 and it was one of the greatest mines at that time. The mine as the first was equipped with steam engine already in the year 1855 and it was connected with the rest of the country via railway line what facilitated coal sales. At the beginning of twentieth century, the mine in question passed through numerous crisis phases that resulted in the employment reductions, and in the year 1932 the mine was even closed for a period of 4 years. However, the mine developed during the Second World War because Germans needed big amounts of good-quality coal. In the period 1940– 1944, the employment was increased threefold up to 3582 workers. After the Second World War, the Charlotta mine was renamed as Rydułtowy mine, which belonged to various structures of Polish mining industry. In the year, the KWK Rydułtowy was joined with Anna mine forming mining plant named as KWK Rydułtowy-Anna.

In the period 1990–1998 a new shaft Leon IV was sunked with diameter 8.5 m, which rarely occured in Polish mining industry. In this period, it was decided that the shaft depth of 1076.2 m would allow development of the new level 1050 m,

which would fully satisfy the mine operational needs. The large resource base of the mine at a depth of less than 1000 m and the need to avoid sub-level mining has become the basis for undertaking another investment consisting in deepening the Leon IV shaft to a depth of 1210.7 allowing development of next exploitation level 1150 m. In the year 2013, design works were started and the process of shaft pipe deepening and extension of two shaft hoists (main and auxiliary) to the depth of 1000 m have been started (https://vimeo.com/321070029).

Flexible-ropes guiding of the shaft hoist cages in the Leon IV were implemented for the first time in Polish hard coal mining industry. Shaft deepening and necessity of extension of shaft hoists to the depth of 1150 m constituted great challenge both for designers and the unit realizing building works. It should be noted that like in each active coal mine, deepening of the active shaft is related with necessity of its continuous and undisturbed operation. In case of such technological restriction, works related with shaft deepening call for special securities, among others leaving of the rock shelf called as natural bottom, or building in the shaft the so-called artificial bottom.

In such cases, transport works in deepened shaft section call for building of auxiliary hoist device with underground hoist machine of special turnstile adapter for personnel transport. Big-diameter hole used for transport, water drainage, and fresh air can greatly facilitate the works related with shaft deepening. However, in such cases, excavation on the level to which the shaft is deepened is needed. The shaft Leon IV can be a good example of application of new technical and technological solutions. Three of such solutions will be discussed in the present study:


#### **2. Single-layer sulfate-proof lining**

In original project of the shaft IV, two-layered lining with hydroinsulating shield made of PE foil was foreseen for the shaft Section 782,0–932,0 characterizing with occurrence of sulfate and magnesium waters. Such linings were commonly used by KOPEX—Shaft Building Company S.A. Sinking technology within the section in question foresees the following works [1]:


Because at this time one of the Polish cement factories produced special Portland cement called as bridge portland cement CP 45(M) marked with symbol CP 45(M)

**81**

**Figure 1.**

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation*

resistant to strong sulfate and magnesium aggression, after research works executed in the AGH University of Science and Technology, modification of the construction and technology of completion of shaft lining section, has been proposed. New project of single-layered lining lied from the top to the bottom following the shaft face advance comprised making of 0.65-m-thick single-layered concrete lining lied in wet system. Due to the dependence of calculated pressure on the shaft lining, its bearing capacity was controlled due to concrete strength with use of two receipts [1] for concrete of class B25 marked as R25/1/2 and for concrete B30 marked as R30/1/4. All concretes were prepared on the basis of Bridge Portland cement CP 45(M). Receipts developed in the AGH University of Science and Technology and verified by laboratory tests guaranteed suitable strength of the concrete and suitable bearing capacity of the lining of targeted thickness, as well as suitable water tightness of the level W8. Concrete lining was made in 4-m-long sections in direction from the top to the bottom. In such technology, waterproofness depends mainly on technological joints between upper (old) section and bottom (new) section. In the project in question, re-sealing of these joints was made first time in Polish shaft building with the use of injection hoses of the type FUKO 2 (**Figure 1**), with former opinion from Higher Mining Office concerning their security due to the presence of methane, after special tests were conducted in Experimental Mine Barbara in Poland. The injection hoses FUKO 2 were mounted to upper section of the shaft with use of metal connectors fitted with screwed joints. After the next lining, the section was concreted, the fissure was filled with a binding mixture on its whole length (see **Figures 1** and **2**) obtaining satisfactory sealing of the neuralgic element

Another issue that was solved during the construction of this shaft was the rock drainage system behind the lining. It is well known that water accumulation behind a waterproof lining is dangerous due to the possibility of high hydrostatic pressures appearing on the casing after joining various aquifers with a shaft.

*Sealing system of the concrete shaft lining with use of hoses FUKO2 [1]. (a) Injection hose FUKO 2. Markings: 1—injection channel Φ = 10 mm; 2—hose core; 3—injection holes; and 4—neoprene ribbons playing role of non-return valves. (b) Housing of hoses in technological joint of sequent sections of the concrete shaft lining. Markings: 1—injection hose FUKO 2; 2—steel pipes; 3—threaded ending for pressure hoses; 4—technological* 

*joint between two concrete lining section; and 5—drainage pipeline for the rock body dewatering.*

*DOI: http://dx.doi.org/10.5772/intechopen.92593*

of the shaft lining.

#### *Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation DOI: http://dx.doi.org/10.5772/intechopen.92593*

*Mining Techniques - Past, Present and Future*

**2. Single-layer sulfate-proof lining**

question foresees the following works [1]:

(currently C12/15 concrete class),

concrete B30 (at present C25/30), and

walls and tight sealing of the 2 mm thick PE foil jacket.

which would fully satisfy the mine operational needs. The large resource base of the mine at a depth of less than 1000 m and the need to avoid sub-level mining has become the basis for undertaking another investment consisting in deepening the Leon IV shaft to a depth of 1210.7 allowing development of next exploitation level 1150 m. In the year 2013, design works were started and the process of shaft pipe deepening and extension of two shaft hoists (main and auxiliary) to the depth of

Flexible-ropes guiding of the shaft hoist cages in the Leon IV were implemented for the first time in Polish hard coal mining industry. Shaft deepening and necessity of extension of shaft hoists to the depth of 1150 m constituted great challenge both for designers and the unit realizing building works. It should be noted that like in each active coal mine, deepening of the active shaft is related with necessity of its continuous and undisturbed operation. In case of such technological restriction, works related with shaft deepening call for special securities, among others leaving of the rock shelf called as natural bottom, or building in the shaft the so-called artificial bottom. In such cases, transport works in deepened shaft section call for building of auxiliary hoist device with underground hoist machine of special turnstile adapter for personnel transport. Big-diameter hole used for transport, water drainage, and fresh air can greatly facilitate the works related with shaft deepening. However, in such cases, excavation on the level to which the shaft is deepened is needed. The shaft Leon IV can be a good example of application of new technical and technological solutions. Three of such solutions will be discussed in the present study:

• single-layer waterproof lining within the Section 782.0–932.0 m,

• shaft deepening technology within the Section 1076.2–1210.7 m, and

• extending of shaft hoists from the level 1000.0 (960 m) to the mining level 1050 m and auxiliary level 1200 m used for the mine water drainage.

In original project of the shaft IV, two-layered lining with hydroinsulating shield made of PE foil was foreseen for the shaft Section 782,0–932,0 characterizing with occurrence of sulfate and magnesium waters. Such linings were commonly used by KOPEX—Shaft Building Company S.A. Sinking technology within the section in

• between ordinates 784.5 and 786.0 m: building of the B15 class concrete lining

• between ordinates 930.5 and 932.0 m: building of shaft brick made of B30 class

• making of the inner layer of the final lining from concrete class B25 and B30 (at present C20/25 and C25/30) wet laid from the bottom up after laying on the

Because at this time one of the Polish cement factories produced special Portland cement called as bridge portland cement CP 45(M) marked with symbol CP 45(M)

• between ordinates 786.0 and 932.0 m: building of preliminary shaft lining (in direction from top to the bottom) in form of 0.56 m thick shaft concrete block wall,

1000 m have been started (https://vimeo.com/321070029).

**80**

resistant to strong sulfate and magnesium aggression, after research works executed in the AGH University of Science and Technology, modification of the construction and technology of completion of shaft lining section, has been proposed. New project of single-layered lining lied from the top to the bottom following the shaft face advance comprised making of 0.65-m-thick single-layered concrete lining lied in wet system. Due to the dependence of calculated pressure on the shaft lining, its bearing capacity was controlled due to concrete strength with use of two receipts [1] for concrete of class B25 marked as R25/1/2 and for concrete B30 marked as R30/1/4. All concretes were prepared on the basis of Bridge Portland cement CP 45(M).

Receipts developed in the AGH University of Science and Technology and verified by laboratory tests guaranteed suitable strength of the concrete and suitable bearing capacity of the lining of targeted thickness, as well as suitable water tightness of the level W8. Concrete lining was made in 4-m-long sections in direction from the top to the bottom. In such technology, waterproofness depends mainly on technological joints between upper (old) section and bottom (new) section. In the project in question, re-sealing of these joints was made first time in Polish shaft building with the use of injection hoses of the type FUKO 2 (**Figure 1**), with former opinion from Higher Mining Office concerning their security due to the presence of methane, after special tests were conducted in Experimental Mine Barbara in Poland. The injection hoses FUKO 2 were mounted to upper section of the shaft with use of metal connectors fitted with screwed joints. After the next lining, the section was concreted, the fissure was filled with a binding mixture on its whole length (see **Figures 1** and **2**) obtaining satisfactory sealing of the neuralgic element of the shaft lining.

Another issue that was solved during the construction of this shaft was the rock drainage system behind the lining. It is well known that water accumulation behind a waterproof lining is dangerous due to the possibility of high hydrostatic pressures appearing on the casing after joining various aquifers with a shaft.

#### **Figure 1.**

*Sealing system of the concrete shaft lining with use of hoses FUKO2 [1]. (a) Injection hose FUKO 2. Markings: 1—injection channel Φ = 10 mm; 2—hose core; 3—injection holes; and 4—neoprene ribbons playing role of non-return valves. (b) Housing of hoses in technological joint of sequent sections of the concrete shaft lining. Markings: 1—injection hose FUKO 2; 2—steel pipes; 3—threaded ending for pressure hoses; 4—technological joint between two concrete lining section; and 5—drainage pipeline for the rock body dewatering.*

**Figure 2.** *Injection system via hoses FUKO 2 and rock body drainage system behind the shaft lining [1].*

This problem was solved by laying vertical drainage 100 mm diameter pipelines on four azimuths along the lining arranged (**Figure 2**).

#### **3. Chosen elements of the shaft Leon IV deepening**

This, about 80-million PLN, investment in the Leon IV shaft deepening by next 140 m resulted from the necessity of the production processes' modification in the Rydułtowy mine [2]. This modification comprised first of all of shortening of the time of personnel transport to the shift face, as well as facilitation of the needed materials' delivery and considerable improvement of the ventilation of this part of the mine.

The investment task related with development of mine infrastructure in Leon IV region comprised the following activities:


Targeted depth of 1210.7 m was reached in August 2016. After completion of the reinforcing of the deepened shaft part in the year 2017, pioneering works in Polish

**83**

**Figure 3.**

*Leon IV shaft profile [2].*

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation*

mining industry works related with elongation of shaft hoists, including untypical and difficult elongation of guiding hoists accompanied by exchange of all leading

The Leon IV shaft was deepened to level 960 m, keeping full exploitation ability.

Works related with shaft deepening were conducted with artificial bottom of

*DOI: http://dx.doi.org/10.5772/intechopen.92593*

ropes, have been executed (**Figure 3**).

**3.1 Technology of the Leon IV shaft deepening**

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation DOI: http://dx.doi.org/10.5772/intechopen.92593*

mining industry works related with elongation of shaft hoists, including untypical and difficult elongation of guiding hoists accompanied by exchange of all leading ropes, have been executed (**Figure 3**).

#### **3.1 Technology of the Leon IV shaft deepening**

*Mining Techniques - Past, Present and Future*

This problem was solved by laying vertical drainage 100 mm diameter pipelines

*Injection system via hoses FUKO 2 and rock body drainage system behind the shaft lining [1].*

This, about 80-million PLN, investment in the Leon IV shaft deepening by next 140 m resulted from the necessity of the production processes' modification in the Rydułtowy mine [2]. This modification comprised first of all of shortening of the time of personnel transport to the shift face, as well as facilitation of the needed materials' delivery and considerable improvement of the ventilation of this part of the mine.

The investment task related with development of mine infrastructure in Leon IV

• making the two-way inlet to pit bottom of exploitation level at the depth of

• making the inlet to single-way pit bottom at the depth of 1200 m destined for

• elongation of mining hoists: main to the level of 1150 m and auxiliary to the

• installation of needed elements of mechanical equipment of the inlets to pit

Targeted depth of 1210.7 m was reached in August 2016. After completion of the reinforcing of the deepened shaft part in the year 2017, pioneering works in Polish

on four azimuths along the lining arranged (**Figure 2**).

**3. Chosen elements of the shaft Leon IV deepening**

region comprised the following activities:

• physical shaft deepening and reinforcing,

needs of the mine main drainage system.

• technical designs,

depth of 1200 m, and

bottoms of both built levels.

150 m,

**Figure 2.**

**82**

The Leon IV shaft was deepened to level 960 m, keeping full exploitation ability. Works related with shaft deepening were conducted with artificial bottom of

**Figure 3.** *Leon IV shaft profile [2].*

special construction of two platforms joined with vertical partition. Mining works were conducted by standard method with use of explosives. Because haulage is the greatest problem in deepened shafts, in the case in question, at first a dike was made on level 1200 m, in order to make great diameter hole of the length of 115 m reaching the shaft bottom before deepening. The hole was located in such manner that its axis was located at a distance of 2.2 m toward east from the shaft axis, which allowed collision-free operation of the basket covering the hole inlet in the shaft face. The 1200 mm diameter hole was nevertheless endangered by the possibility of rock blockage. In order to remove the jam in the hole, a rope of 25 mm diameter with conveyor scrappers was installed in the hole. The rope vertical movements stimulated by low-speed winches KUBA-5 installed in ditches on the level 1076 and 1200 m, (see **Figure 4**) provoke fall down and the hole clearance.

Transport in the whole deepening process was handled by special devices located in ditch on level 1076 m (**Figure 4**):


In the ditch, at the level 1200 m, low-speed windlass KUBA-5 with track wheel for the hole clearing rope (see **Figure 4**) was installed. Single–layered lining of C30/37 concrete lied in wet system with use of steel moveable formwork of the height 2.15 m, has been applied. Calculated and consulted with the Investor lining has thickness from 0.5 to 0.6 m. With respect to expected small and ephemeral water inflows into the shaft, no special waterproof precautions were designed [4].

#### **3.2 Start-up of levels 1150 and 1200 m**

Thanks to the Ruch Rydułtowy investment, it will allow the exploitation from coal seams No 713/1–2 and 712/1–2, which belong to the most promising mining assets within mining areas belonging to this part of Rybnik Mining District. Development of this part of the deposit will allow building the new level at the depth of 1150 m (https://vimeo.com/320940852). Two-way inlet is equipped with full set of the wheel transport handling, with special platform for material reloading from wheel into lifted gondola transport. The inlet performances are as follows:


The pit-bottom geometry with use of 3D visualization is shown in **Figure 5**. Universality is characteristic feature of the shaft pit–bottom 1150 m—main transport level (https://vimeo.com/333420960). Within this level there is a possibility

**85**

**Figure 5.**

**Figure 4.**

*steel-concrete pit-bottom lining.*

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation*

of using three shaft hoists, what will considerably accelerate process of material lifting, as well as it will allow fast and fluent personnel transport. Using transport platforms forced equipping shaft pit-bottom basement with devices and machines needed for pushing mine trucks into large-size mining cages, as well as into standard cages. The deepening of the Leon IV shaft was also used to reorganize water management in this area. For this purpose, excavations needed for the main water drainage handling were localized in one-way pit-bottom at the level 1200 m, and the elongation of this level to auxiliary hoist was necessary, and it was requalified

*3D model of the pit-bottom—level 1150 m [3]. (a) General view; and (b) and (c) 3D model of the* 

*Distribution of devices during Leon IV shaft deepening at levels 1076 and 1200 m [2].*

*DOI: http://dx.doi.org/10.5772/intechopen.92593*

from auxiliary hoist into "small" hoist.

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation DOI: http://dx.doi.org/10.5772/intechopen.92593*

*Mining Techniques - Past, Present and Future*

located in ditch on level 1076 m (**Figure 4**):

• windlass KUBA-5 for rope handling,

**3.2 Start-up of levels 1150 and 1200 m**

• excavation founding depth: 1143.7 m,

• width: W side—8.11 m and E side—7.0 m, and

• height: 7.3 m,

• basement depth: 2.30 m.

• hoist machine B-1500 for bucket handling,

special construction of two platforms joined with vertical partition. Mining works were conducted by standard method with use of explosives. Because haulage is the greatest problem in deepened shafts, in the case in question, at first a dike was made on level 1200 m, in order to make great diameter hole of the length of 115 m reaching the shaft bottom before deepening. The hole was located in such manner that its axis was located at a distance of 2.2 m toward east from the shaft axis, which allowed collision-free operation of the basket covering the hole inlet in the shaft face. The 1200 mm diameter hole was nevertheless endangered by the possibility of rock blockage. In order to remove the jam in the hole, a rope of 25 mm diameter with conveyor scrappers was installed in the hole. The rope vertical movements stimulated by low-speed winches KUBA-5 installed in ditches on the level 1076 and

Transport in the whole deepening process was handled by special devices

• two low-speed winches KUBA-10 for adjustable formwork handling,

• windlass KCH-9 for basket hanging protecting hole in shaft face, and

• supporting construction for assemblage of wheels during shaft deepening.

In the ditch, at the level 1200 m, low-speed windlass KUBA-5 with track wheel for the hole clearing rope (see **Figure 4**) was installed. Single–layered lining of C30/37 concrete lied in wet system with use of steel moveable formwork of the height 2.15 m, has been applied. Calculated and consulted with the Investor lining has thickness from 0.5 to 0.6 m. With respect to expected small and ephemeral water inflows into the shaft, no special waterproof precautions were designed [4].

Thanks to the Ruch Rydułtowy investment, it will allow the exploitation from coal seams No 713/1–2 and 712/1–2, which belong to the most promising mining assets within mining areas belonging to this part of Rybnik Mining District. Development of this part of the deposit will allow building the new level at the depth of 1150 m (https://vimeo.com/320940852). Two-way inlet is equipped with full set of the wheel transport handling, with special platform for material reloading from wheel into lifted gondola transport. The inlet performances are as follows:

The pit-bottom geometry with use of 3D visualization is shown in **Figure 5**. Universality is characteristic feature of the shaft pit–bottom 1150 m—main transport level (https://vimeo.com/333420960). Within this level there is a possibility

1200 m, (see **Figure 4**) provoke fall down and the hole clearance.

**84**

of using three shaft hoists, what will considerably accelerate process of material lifting, as well as it will allow fast and fluent personnel transport. Using transport platforms forced equipping shaft pit-bottom basement with devices and machines needed for pushing mine trucks into large-size mining cages, as well as into standard cages. The deepening of the Leon IV shaft was also used to reorganize water management in this area. For this purpose, excavations needed for the main water drainage handling were localized in one-way pit-bottom at the level 1200 m, and the elongation of this level to auxiliary hoist was necessary, and it was requalified from auxiliary hoist into "small" hoist.

**Figure 4.** *Distribution of devices during Leon IV shaft deepening at levels 1076 and 1200 m [2].*

#### **Figure 5.**

*3D model of the pit-bottom—level 1150 m [3]. (a) General view; and (b) and (c) 3D model of the steel-concrete pit-bottom lining.*

Geometry of the shaft inlet at level 1200 m:


The shaft inlet has anchor-concrete-steel lining and is equipped with level guidance construction with oscillatory platform in the inlet basement.

#### **3.3 Furnishing of the shaft Leon IV**

As the main transport shaft, the shaft Leon IV is equipped with three compartments: one for main hoist with large-size cage, second for ordinary three-deck cage, and auxiliary hoist. The shaft cages are suspended on two 48-mm-diameter rope carriers driven by drive wheel Koeppe. In order to balance masses of rope carriers, two equalizing ropes of diameter Φ = 53 mm are installed.

Shaft Leon IV is the first shaft in Polish hard coal mining industry, in which flexible guiding of shaft cages or skips has been extended. Guiding and defender ropes are suspended on wedge-shaped spreader beams located over beams of the shaft tower. The guiding and defender ropes hang down freely and are tensed by the attached weights of such mass that each 100 m of the shaft depth corresponds to tensing power of the value at least 8 kN. The guiding and defender ropes are mounted in special baskets located below lower guiding rope

#### **Figure 6.**

*Model of the Leon IV shaft furniture after deepening to the depth of 1210 m [2]. Markings: 1—level guidance at level 1150 m; 2—breaking system of the main shaft hoist; 3—return station of the equalizing ropes; 4—control platform of the equalizing ropes; 5—control platforms of guiding rope weights; 6—level guidance at level 1200 m; and 7—furniture of the Leon IV shaft sump.*

**87**

**Figure 7.**

*Elements of the elastic guiding of shaft Leon IV.*

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation*

frame. In case of the Leon IV shaft, the shaft furniture consists of the following

Profits resulting from application of rope guiding of hoist cages or skips are

• guiding of the cages ore skips in the shaft is soft, without of tremors and side

• quiet run cages ore skips results in elongation of the rope carrier, and

• ventilation resistances are almost 10 times lower than in case of shafts with

Elastic guiding of the shaft cages comprises 12 guiding ropes, 4 defender ropes, and 3 rope carriers and equalizing ropes between large-size cage and three-deck

*DOI: http://dx.doi.org/10.5772/intechopen.92593*

elements: (**Figures 6** and **7**).

great. The profits are as follows:

• low cost of used materials,

• short assembling time in new shaft,

• possibility of fast shaft hoist operation,

cage (**Figure 7**).

• easy handling,

• long durability,

rigid guides.

hits,

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation DOI: http://dx.doi.org/10.5772/intechopen.92593*

frame. In case of the Leon IV shaft, the shaft furniture consists of the following elements: (**Figures 6** and **7**).

Elastic guiding of the shaft cages comprises 12 guiding ropes, 4 defender ropes, and 3 rope carriers and equalizing ropes between large-size cage and three-deck cage (**Figure 7**).

Profits resulting from application of rope guiding of hoist cages or skips are great. The profits are as follows:


*Mining Techniques - Past, Present and Future*

**3.3 Furnishing of the shaft Leon IV**

• height: 4.2 m, and

• width: 6.1 m.

Geometry of the shaft inlet at level 1200 m:

• depth of the excavation founding: 1195.7 m,

ance construction with oscillatory platform in the inlet basement.

two equalizing ropes of diameter Φ = 53 mm are installed.

The shaft inlet has anchor-concrete-steel lining and is equipped with level guid-

As the main transport shaft, the shaft Leon IV is equipped with three compartments: one for main hoist with large-size cage, second for ordinary three-deck cage, and auxiliary hoist. The shaft cages are suspended on two 48-mm-diameter rope carriers driven by drive wheel Koeppe. In order to balance masses of rope carriers,

Shaft Leon IV is the first shaft in Polish hard coal mining industry, in which flexible guiding of shaft cages or skips has been extended. Guiding and defender ropes are suspended on wedge-shaped spreader beams located over beams of the shaft tower. The guiding and defender ropes hang down freely and are tensed by the attached weights of such mass that each 100 m of the shaft depth corresponds to tensing power of the value at least 8 kN. The guiding and defender ropes are mounted in special baskets located below lower guiding rope

*Model of the Leon IV shaft furniture after deepening to the depth of 1210 m [2]. Markings: 1—level guidance at level 1150 m; 2—breaking system of the main shaft hoist; 3—return station of the equalizing ropes; 4—control platform of the equalizing ropes; 5—control platforms of guiding rope weights; 6—level guidance at level* 

**86**

**Figure 6.**

*1200 m; and 7—furniture of the Leon IV shaft sump.*


**Figure 7.** *Elements of the elastic guiding of shaft Leon IV.*

Linear guiding of cages has also some disadvantages, like:


The other element of the shaft furniture is related with main hoist cages' braking system localized under level 1150 m and in the shaft tower. This system is composed of thickened wood guides. The other elements of the shaft furniture comprise:


Necessity of equipping the flexible guidance with special corner guiding for shaft vessels is essential element of guiding on individual levels. In case of the shaft Leon IV it refers to levels 800, 1067, 1150, and 1200 m.

#### **3.4 Elongation of mine shaft hoists**

Shaft deepening is strictly related with necessity of mine shaft hoists. In case of the Leon IV shaft, in construction of elastic guiding of hoist cages to level 1067 m, the ropes longer than the exploited shaft were applied. Excess of ropes was stored on special drums located in pit-bottom of the level 1076 m. Thus elongation of these ropes required only suitable control of the rope destruction degree and then lowering them to the level 1150 m. Works related with elongation of the mine

**89**

**Figure 8.**

*Visualization of sheave wheel location on hoist tower [2].*

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation*

shaft hoist were started from auxiliary hoist and then lowering two guiding ropes of diameter Ø = 32 mm to level 1200 m. Weight of the rope of diameter 32 mm amounts for around 5.3 ton, whereas for ropes of diameter Ø = 54 mm, this weight reaches value over 25 ton. Thus, every rope maneuvers, that is, their raising or requires execution of assembling works with use of special hoist machine having high lifting capacity, as well as special guiding wheels mounted on special platform. Low-speed lifting machine EWP-35 was used for rope lifting and lowering. After making welded clamps and taking rope weight by lifting machine, disassembling was made in order to check condition of ropes by suitable expert. After acceptance of the ropes for exploitation, they were lowered to level 1200 m, where weighting baskets were mounted. Next stage of elongation of the shaft hoist comprised elongation of guiding and bumper ropes of main hoist, that is, unwrapping rope reserve accumulated at the level 1067 m. After installation of additional constructions of rope wheels on the platform on lower rope wheel of the Leon IV shaft, that is, directing technological rope of Ø = 40 mm in place of wedge-shaped spreader beam, stage of basal works related with taking and lowering ropes in targeted place have been started. When the welded clamps were installed and the weight was taken by low-speed lifting machine (40 ton), the rope was lifted to level of the foundation in order to be checked by the expert, and then the rope was lowered again to the level of weights control platform. After completion of the operations of the first rope, the guiding wheel was relocated on the platform in such a manner that operations of the next ropes would be possible. This operation was repeated eight times for guiding lines of the main hoist cages, and four times for bumper ropes located between both vessels of the same hoist. Scheme of devices needed for hoist elongation and visualization of the guiding wheels on the

Technology of the rope carriers and equalizing ropes in the shaft Leon IV is similar to standard technologies of ropes operated in mine shafts. In this case, at first, some preparatory works related with construction of foundation of the lift EPR-1000 and installation of sheave wheels have been executed in the shaft foundation. Carrying ropes were lowered to the shaft after placing shaft cages on special

After relocation of the great-size cage to the level 1150 m and installation of spreader beams, the equalizing ropes were elongated. In this case, the rising

platform and taking the weight of ropes by portable lift EPR-1000.

*DOI: http://dx.doi.org/10.5772/intechopen.92593*

tower are shown in **Figure 8**.

#### *Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation DOI: http://dx.doi.org/10.5772/intechopen.92593*

*Mining Techniques - Past, Present and Future*

different directions is needed.

Linear guiding of cages has also some disadvantages, like:

shaft hoist devices in numerous transport sections,

• rope tensioning devices require a properly managed sump,

• rope carrier must be made of non-rotating rope, and

• platform of the return station of equalizing rope,

• control platforms of guiding and bumper ropes, and

Leon IV it refers to levels 800, 1067, 1150, and 1200 m.

• platform of equalizing ropes control system,

• protective platform.

**3.4 Elongation of mine shaft hoists**

• serious difficulties in exploitation result—rock body movements,

• possibility of transverse movements, which are vertical to the running direction, which calls for longer operation intervals than in case of rigid guides,

• need of great size shaft diameters and difficulties related with relocation of

• tensions of guiding ropes reaching values of 2000 kN causing additional loading of the shaft tower and application of suitable construction needed.

• in case of application of two-cage hoists, the application of additional ropes called as "defender ropes" protecting against collisions of vessels moving in two

• the main advantages of the application of such type of guiding result from the

• the main benefits of using this type of guide arise from the fact that the shaft

The other element of the shaft furniture is related with main hoist cages' braking system localized under level 1150 m and in the shaft tower. This system is composed of thickened wood guides. The other elements of the shaft furniture comprise:

fact that the use of tens shaft fastening frames is not necessary.

does not require the installation of dozens of rigid guides frames.

• positioning frame of weighs of guiding ropes and bumper ropes,

Necessity of equipping the flexible guidance with special corner guiding for shaft vessels is essential element of guiding on individual levels. In case of the shaft

Shaft deepening is strictly related with necessity of mine shaft hoists. In case of the Leon IV shaft, in construction of elastic guiding of hoist cages to level 1067 m, the ropes longer than the exploited shaft were applied. Excess of ropes was stored on special drums located in pit-bottom of the level 1076 m. Thus elongation of these ropes required only suitable control of the rope destruction degree and then lowering them to the level 1150 m. Works related with elongation of the mine

**88**

shaft hoist were started from auxiliary hoist and then lowering two guiding ropes of diameter Ø = 32 mm to level 1200 m. Weight of the rope of diameter 32 mm amounts for around 5.3 ton, whereas for ropes of diameter Ø = 54 mm, this weight reaches value over 25 ton. Thus, every rope maneuvers, that is, their raising or requires execution of assembling works with use of special hoist machine having high lifting capacity, as well as special guiding wheels mounted on special platform. Low-speed lifting machine EWP-35 was used for rope lifting and lowering. After making welded clamps and taking rope weight by lifting machine, disassembling was made in order to check condition of ropes by suitable expert. After acceptance of the ropes for exploitation, they were lowered to level 1200 m, where weighting baskets were mounted. Next stage of elongation of the shaft hoist comprised elongation of guiding and bumper ropes of main hoist, that is, unwrapping rope reserve accumulated at the level 1067 m. After installation of additional constructions of rope wheels on the platform on lower rope wheel of the Leon IV shaft, that is, directing technological rope of Ø = 40 mm in place of wedge-shaped spreader beam, stage of basal works related with taking and lowering ropes in targeted place have been started. When the welded clamps were installed and the weight was taken by low-speed lifting machine (40 ton), the rope was lifted to level of the foundation in order to be checked by the expert, and then the rope was lowered again to the level of weights control platform. After completion of the operations of the first rope, the guiding wheel was relocated on the platform in such a manner that operations of the next ropes would be possible. This operation was repeated eight times for guiding lines of the main hoist cages, and four times for bumper ropes located between both vessels of the same hoist. Scheme of devices needed for hoist elongation and visualization of the guiding wheels on the tower are shown in **Figure 8**.

Technology of the rope carriers and equalizing ropes in the shaft Leon IV is similar to standard technologies of ropes operated in mine shafts. In this case, at first, some preparatory works related with construction of foundation of the lift EPR-1000 and installation of sheave wheels have been executed in the shaft foundation. Carrying ropes were lowered to the shaft after placing shaft cages on special platform and taking the weight of ropes by portable lift EPR-1000.

After relocation of the great-size cage to the level 1150 m and installation of spreader beams, the equalizing ropes were elongated. In this case, the rising

**Figure 8.** *Visualization of sheave wheel location on hoist tower [2].*

large-size cage pulled new equalizing ropes. After relocation of three floor cages to the level 1150 m, construction of equalizing ropes under foot of three floor cage was completed.

#### **3.5 Stabilization of the cage at the large-size Leon IV shaft hoisting system**

Adjustable guiding system replaced the chairing mechanism of a rope guided cage at level 960 of "Leon IV" shaft. The mechanical part of the system is included in the project (**Figure 9**), developed to solve the issue of leading the cage through level 960. The most important factor taken into consideration was safety. With chairing, it could be ensured only at the expense of significant lengthening of the duration of a single cage ride from the ground level to level 1150 m and vice versa, since the velocity of the cage had to be reduced from 10 to 0.5 m/s, already 100 m before level 960 m. The changes between systems included replacing the angular guides with adjustable guides (including two pairs of upper and two pairs of lower guides) and the main support with four permanent frames, fixed to the shaft lining using additional girders, structurally independent from the construction of the chairing. In this arrangement, the adjustable guides can be switched between idle mode and working mode. The motion is restricted by:


Two pairs of lower adjustable guides are powered by a single hydraulic cylinder each, one end articulated to the guiding frame, the other to the guides itself. Each pair of upper guides is powered by two hydraulic cylinders, articulated with lower end to the frame, and upper end to the joint.

**91**

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation*

During the hoisting operation (whether it is from the ground level to level 1150 m or in the opposite direction), each pair of the adjustable guides stays in idle

This setup ensures safe movement of the cage through level 960 with velocity of 10 meters per second. Idle mode is also used when the cage moves between level 950 m and 1150 m (both directions). Setting to working mode, decided by the signalman at level 960, takes place once the cage stops at level 960. During the cycle each of the upper ends of the lower adjustable guides is inserted between pads of the emergency braking slide guide, attached to the face of the bottom deck. The upper ends of the upper adjustable guides are inserted between pads of the emergency breaking slide guide attached to the face of the cage's head. The work mode setup of the bottom and top adjustable guides is shown in

Layout of the entire system is shown in **Figure 11**. The adjustable guides consist of 180 x 260 mm box beams made of C260 C-profiles. Frames made of HEB 260 are

Static analysis was performed for the constructions. Assumptions and results are

= 120.92 [

\_ *kN*

62.87= 7.63 (2)

*<sup>m</sup>* ] (1)

*<sup>m</sup>* ] (3)

attached to two technological beams with M24 Hex bolts, class 8.8.

1.654 [*m*]

\_ 480

The assumed load rounded to the value of 121 kN/m

• Force from the hydraulic cylinder: 139.2 [kN] (**Figure 12**)

1.654 [*m*]

\_ 480

The assumed load rounded to the value of 121 kN/m

• Force from the hydraulic cylinder: 139.2 [kN] (**Figure 13**)

= 120.92 [

\_ *kN*

75.09= 6.392 (4)

*DOI: http://dx.doi.org/10.5772/intechopen.92593*

mode (**Figure 9**).

**Figure 10**.

shown in the schemes below:

*Known variables*:

• Continuous load:

• Factor of safety

*Known variables*:

• Continuous load:

• Factor of safety

**Static analysis—lower frame**

**Static analysis—upper frame**

• Transportation unit load: 200 [kN]

*<sup>q</sup>* = \_200 [*kN*]

• Transportation unit load: 200 [kN]

*<sup>q</sup>* = \_200 [*kN*]

**Figure 9.** *System in idle mode.*

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation DOI: http://dx.doi.org/10.5772/intechopen.92593*

During the hoisting operation (whether it is from the ground level to level 1150 m or in the opposite direction), each pair of the adjustable guides stays in idle mode (**Figure 9**).

This setup ensures safe movement of the cage through level 960 with velocity of 10 meters per second. Idle mode is also used when the cage moves between level 950 m and 1150 m (both directions). Setting to working mode, decided by the signalman at level 960, takes place once the cage stops at level 960. During the cycle each of the upper ends of the lower adjustable guides is inserted between pads of the emergency braking slide guide, attached to the face of the bottom deck. The upper ends of the upper adjustable guides are inserted between pads of the emergency breaking slide guide attached to the face of the cage's head. The work mode setup of the bottom and top adjustable guides is shown in **Figure 10**.

Layout of the entire system is shown in **Figure 11**. The adjustable guides consist of 180 x 260 mm box beams made of C260 C-profiles. Frames made of HEB 260 are attached to two technological beams with M24 Hex bolts, class 8.8.

Static analysis was performed for the constructions. Assumptions and results are shown in the schemes below:

### **Static analysis—lower frame**

*Known variables*:

*Mining Techniques - Past, Present and Future*

completed.

motion is restricted by:

guides,

each pair of lower adjustable guides,

end to the frame, and upper end to the joint.

large-size cage pulled new equalizing ropes. After relocation of three floor cages to the level 1150 m, construction of equalizing ropes under foot of three floor cage was

**3.5 Stabilization of the cage at the large-size Leon IV shaft hoisting system**

Adjustable guiding system replaced the chairing mechanism of a rope guided cage at level 960 of "Leon IV" shaft. The mechanical part of the system is included in the project (**Figure 9**), developed to solve the issue of leading the cage through level 960. The most important factor taken into consideration was safety. With chairing, it could be ensured only at the expense of significant lengthening of the duration of a single cage ride from the ground level to level 1150 m and vice versa, since the velocity of the cage had to be reduced from 10 to 0.5 m/s, already 100 m before level 960 m. The changes between systems included replacing the angular guides with adjustable guides (including two pairs of upper and two pairs of lower guides) and the main support with four permanent frames, fixed to the shaft lining using additional girders, structurally independent from the construction of the chairing. In this arrangement, the adjustable guides can be switched between idle mode and working mode. The

• Part of a respective frame, known as roadway, serving as adjusting track for

• Two articulated links (upper and lower), for each pair of upper adjustable

Two pairs of lower adjustable guides are powered by a single hydraulic cylinder each, one end articulated to the guiding frame, the other to the guides itself. Each pair of upper guides is powered by two hydraulic cylinders, articulated with lower

**90**

**Figure 9.** *System in idle mode.*


$$q = \frac{200 \text{ [}kN\text{]}}{1.654 \text{ [}m\text{]}} = 120.92 \left[\frac{kN}{m}\right] \tag{1}$$

The assumed load rounded to the value of 121 kN/m

• Factor of safety

$$\frac{480}{62.87} = 7.63\tag{2}$$

• Force from the hydraulic cylinder: 139.2 [kN] (**Figure 12**)

#### **Static analysis—upper frame** *Known variables*:


$$q = \frac{200 \text{ [}kN\text{]}}{1.654 \text{ [}m\text{]}} = 120.92 \left[\frac{kN}{m}\right] \tag{3}$$

The assumed load rounded to the value of 121 kN/m

• Factor of safety

$$\frac{480}{75.09} = 6.392\tag{4}$$

• Force from the hydraulic cylinder: 139.2 [kN] (**Figure 13**)

**93**

guides.

*Static analysis of the upper frame.*

*Static analysis for the bottom frame.*

**Figure 13.**

**Figure 12.**

during the experiments.

lower deck,

operations. The cycles were divided in two stages:

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation*

• maximum stresses induced in girders of the cage during experimental loading and unloading cycles, being the result of absorbing forces from the adjustable

The force measurements were conducted accordingly to the previous ones, carried out on exactly the same cage in July 2018, also concerning the forces absorbed the cage during the experimental loading and unloading cycles, but forces from angular guides were replaced with forces from adjustable guides. Moreover, the scope did not include measuring the stresses induced in girders as a result of absorbing forces from angular guides, as these guides were also stabilizing the middle deck

The reason for adding stress measurement was the necessity to properly assess whether resigning from stabilization of the middle deck of the cage in this arrangement may be considered a viable assumption. Theoretical analysis based on calculation model from [5] demonstrated such possibility; however, empirical verification was deemed vital. It was implemented by equipping the main measurement unit with two external modules attached to the middle deck for the time of measurement

• Stage one, covering the first six experimental loading and unloading cycles, concerned measuring the forces at the bottom deck of the cage as well as stresses induced in girders on the distance between the middle deck and the

*DOI: http://dx.doi.org/10.5772/intechopen.92593*

**Figure 10.** *System in working mode.*

#### **Figure 11.**

*Layout of the adjustable guiding (gray color).*

#### **Measurements**

The main purpose of measurements was to determine:

• maximum forces applied to the head and the bottom deck of the cage during experimental cycles of loading and unloading the heaviest transportation unit approved for this type of transportation, performed at level 960,

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation DOI: http://dx.doi.org/10.5772/intechopen.92593*

**Figure 12.** *Static analysis for the bottom frame.*

*Mining Techniques - Past, Present and Future*

**92**

**Measurements**

*Layout of the adjustable guiding (gray color).*

**Figure 11.**

**Figure 10.**

*System in working mode.*

The main purpose of measurements was to determine:

• maximum forces applied to the head and the bottom deck of the cage during experimental cycles of loading and unloading the heaviest transportation unit

approved for this type of transportation, performed at level 960,

**Figure 13.** *Static analysis of the upper frame.*

• maximum stresses induced in girders of the cage during experimental loading and unloading cycles, being the result of absorbing forces from the adjustable guides.

The force measurements were conducted accordingly to the previous ones, carried out on exactly the same cage in July 2018, also concerning the forces absorbed the cage during the experimental loading and unloading cycles, but forces from angular guides were replaced with forces from adjustable guides. Moreover, the scope did not include measuring the stresses induced in girders as a result of absorbing forces from angular guides, as these guides were also stabilizing the middle deck during the experiments.

The reason for adding stress measurement was the necessity to properly assess whether resigning from stabilization of the middle deck of the cage in this arrangement may be considered a viable assumption. Theoretical analysis based on calculation model from [5] demonstrated such possibility; however, empirical verification was deemed vital. It was implemented by equipping the main measurement unit with two external modules attached to the middle deck for the time of measurement operations. The cycles were divided in two stages:

• Stage one, covering the first six experimental loading and unloading cycles, concerned measuring the forces at the bottom deck of the cage as well as stresses induced in girders on the distance between the middle deck and the lower deck,

• Stage two, covering the following six experimental cycles, concerned measuring the forces at the top deck of the cage as well as stresses induced in girders on the distance between the middle deck and the head.

#### **4. Conclusions**

Technical problems related with the shaft Leon IV sinking and deepening presented in this chapter can be considered as an example of continuous innovation and development of the shaft building technology. Although nowadays shafts are rarely deepened, high level of modern technique and mechanization of both sinking and equipping the shafts indicate potential possibilities of further development of this building branch.

Application of new bridge cements M45 and modification of the philosophy of assuring the shaft lining tightness even during building the shaft Leon IV allowed implementation of very profitable replacement of multilayered lining with singlelayer lining, which is less time-consuming and much cheaper.

Shaft deepening during its exploitation was possible only in result of application of modern construction of "artificial bottom," which tightly separated the shaft from the area of works conducted by the company named as Shaft Building Company S.A. (PBSz).

Application of elastic system in the shaft Leon IV in hard coal mining industry and use of much more longer ropes and storage of the rope surplus on the level 1078 m can be classified as the uniquely far-sighted project. This in turn allowed implementation of much more easy technologies of shaft hoists elongation.

Designed by Shaft Sinking Company, elongation of the shaft hoists, which was realized in possibly shortest stoppage of the shaft operation, was a pioneer and innovative venture. Total work comprised elongation of 20 ropes.

Modernization of the shaft Leon IV was a key element of the restructuring plan and development of the joint-venture mine ROW gathering mines: Jankowice, Chwałowice, Marcel, and Rydułtowy. Elongation of shaft hoists, development of main transport horizon at the level 1150 m, and development of the main drainage system at level 1200 m will allow considerable shortening of the time of personnel transport to exploitation excavations, which is related with considerable improvement and elongation of the personnel working time, that is, improvement of financial results of mine operation and whole mine ROW.

Conducted analysis and research confirm that the mechanical system, its attachment to the shaft lining and particular structural elements of the tested solution are correct and fulfill the conditions defined in the Decree of the Minister of Energy from November 23, 2016 concerning the detailed requirements of operating in underground mines (Journal of Laws of the Republic of Poland 2017, item 1118), as well as in the technical standard PN-G-46227: 2002—Mining shafts. Shaft equipment. Requirements.

Changes introduced at level 960 m (1000 m), made according to Annex No. 2, included removing the corner guides of the cage. The additional space obtained between the chairing elements ensures the safe passage of the cage through level 960 with the set velocity of 10 m/s, assuming the requirement of § 545 of the Decree of the Minister of Energy is met. After analyzing the results of measurements of the forces from bottom deck and the head of the cage absorbed by adjustable guides during loading and unloading cycles, it can be stated that replacing the corner guides with adjustable guides does not violate this approval.

**95**

**Author details**

Paweł Kamiński

AGH University of Science and Technology, Kraków, Poland

© 2020 The Author(s). Licensee IntechOpen. This chapter is distributed under the terms of the Creative Commons Attribution License (http://creativecommons.org/licenses/ by/3.0), which permits unrestricted use, distribution, and reproduction in any medium,

\*Address all correspondence to: pkamin@agh.edu.pl

provided the original work is properly cited.

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation*

*DOI: http://dx.doi.org/10.5772/intechopen.92593*

*Polish Experience in Shaft Deepening and Mining Shaft Hoist Elongation DOI: http://dx.doi.org/10.5772/intechopen.92593*

*Mining Techniques - Past, Present and Future*

**4. Conclusions**

this building branch.

Company S.A. (PBSz).

ment. Requirements.

• Stage two, covering the following six experimental cycles, concerned measuring the forces at the top deck of the cage as well as stresses induced in girders

Technical problems related with the shaft Leon IV sinking and deepening presented in this chapter can be considered as an example of continuous innovation and development of the shaft building technology. Although nowadays shafts are rarely deepened, high level of modern technique and mechanization of both sinking and equipping the shafts indicate potential possibilities of further development of

Application of new bridge cements M45 and modification of the philosophy of assuring the shaft lining tightness even during building the shaft Leon IV allowed implementation of very profitable replacement of multilayered lining with single-

Shaft deepening during its exploitation was possible only in result of application of modern construction of "artificial bottom," which tightly separated the shaft from the area of works conducted by the company named as Shaft Building

Application of elastic system in the shaft Leon IV in hard coal mining industry and use of much more longer ropes and storage of the rope surplus on the level 1078 m can be classified as the uniquely far-sighted project. This in turn allowed implementation of much more easy technologies of shaft hoists elongation.

Designed by Shaft Sinking Company, elongation of the shaft hoists, which was realized in possibly shortest stoppage of the shaft operation, was a pioneer and

Modernization of the shaft Leon IV was a key element of the restructuring plan

Conducted analysis and research confirm that the mechanical system, its attachment to the shaft lining and particular structural elements of the tested solution are correct and fulfill the conditions defined in the Decree of the Minister of Energy from November 23, 2016 concerning the detailed requirements of operating in underground mines (Journal of Laws of the Republic of Poland 2017, item 1118), as well as in the technical standard PN-G-46227: 2002—Mining shafts. Shaft equip-

Changes introduced at level 960 m (1000 m), made according to Annex No. 2, included removing the corner guides of the cage. The additional space obtained between the chairing elements ensures the safe passage of the cage through level 960 with the set velocity of 10 m/s, assuming the requirement of § 545 of the Decree of the Minister of Energy is met. After analyzing the results of measurements of the forces from bottom deck and the head of the cage absorbed by adjustable guides during loading and unloading cycles, it can be stated that replacing the

corner guides with adjustable guides does not violate this approval.

and development of the joint-venture mine ROW gathering mines: Jankowice, Chwałowice, Marcel, and Rydułtowy. Elongation of shaft hoists, development of main transport horizon at the level 1150 m, and development of the main drainage system at level 1200 m will allow considerable shortening of the time of personnel transport to exploitation excavations, which is related with considerable improvement and elongation of the personnel working time, that is, improvement of

on the distance between the middle deck and the head.

layer lining, which is less time-consuming and much cheaper.

innovative venture. Total work comprised elongation of 20 ropes.

financial results of mine operation and whole mine ROW.

**94**

#### **Author details**

Paweł Kamiński AGH University of Science and Technology, Kraków, Poland

\*Address all correspondence to: pkamin@agh.edu.pl

© 2020 The Author(s). Licensee IntechOpen. This chapter is distributed under the terms of the Creative Commons Attribution License (http://creativecommons.org/licenses/ by/3.0), which permits unrestricted use, distribution, and reproduction in any medium, provided the original work is properly cited.

#### **References**

[1] Kostrz J, Olszewski J, Czaja P, Deja J, Witosiński J. Zastosowanie betonów odpornych na silną agresję siarczanową i magnezową w budownictwie podziemnym. Budownictwo Górnicze i Tunelowe. 2000;**3**:4-11. ISSN: 1234-5342

[2] Olszewski J, Czaja P, Bulenda P, Kamiński P. Pogłębianie oraz wydłużanie górniczych wyciągów szybowych—szybu Leon IV w kopalni KWK ROW—Ruch Rydułtowy. Przegląd Górniczy. 2018;**74**(8):7-16. ISSN: 0033-216X

[3] Czaja P, Kamiński P. Wybrane zagadnienia techniki i technologii głębienia szybów. Kraków: Szkoła Eksploatacji Podziemnej. 2016. p. 161. ISBN: 978-83-927920-9-3

[4] Kicki J, Sobczyk EJ, Kamiński P. Vertical and decline shaft sinking— Good practices in technique and technology. In: International Mining Forum 2015; 23-27 February 2015; Cracow, Poland. Leiden: CRC Press/Balkema; 2015. p. 197. ISBN: 978-1-138-02820-3

[5] Płachno M. Mathematical model of transverse vibration of a high-capacity mining skip due misalignment of the guiding tracks in the hoisting shaft. Archives of Mining Sciences (Archiwum Górnictwa). Wyd. Polska Akademia Nauk, Komitet Górnictwa, Kraków. 2018;**63**(1):5. DOI: 10.24425/118882

**97**

**Chapter 6**

Sinking

**Abstract**

**1. Introduction**

Polish Experiences in Handling

*Piotr Czaja, Paweł Kamiński and Artur Dyczko*

shaft to protect the final lining from damage.

**Keywords:** mining shaft, water hazard, grouting, dewatering

Water Hazards during Mine Shaft

The geological structure of most Polish mining regions is rich in groundwater, making shaft sinking difficult. In recent years, more than a dozen shafts, some almost 700 m deep, have been sunk in Poland using various methods of water hazard elimination. The vast majority of shafts that pass through aquifer formations have been sunk using artificial rock freezing, waterproof tubing, and concrete lining. Generally, this system has proven to be very effective. However, there have been cases of complications during sinking, including occasional flooding. This paper presents two cases of highly problematic flooding in shaft sunk using the freezing method, both leading to considerable construction delays and a significant increase in shaft sinking costs. The first case involved water inflow into the bottom section of the R-XI shaft at KGHM with rocks near the melting point of ice. In the other case, problems occurred passing through an Albian layer in the S. 1.3 shaft sunk for the Lubelski Węgiel Bogdanka S.A. mining corporation, where the freezing process was carried out while it was necessary to heat the rocks in the upper part of the

Exposing deep-seated mineral deposits requires the construction of new shafts.

In Poland, where usable minerals are usually covered by thick layers of heavily waterlogged overburden, the construction of new shafts poses extraordinary difficulties. New shafts continue to be designed and constructed in quite challenging hydrogeological conditions in Poland, as well as in other countries worldwide. Hence, it would be fruitful to look at some Polish experiences in coping with this extremely difficult hydrogeology while mining deposits of both hard-coal and nonferrous metal ores. A range of detailed examples of how to eliminate such water hazards has been provided elsewhere [1–3]. Over the last three decades, Poland has seen at least several cases involving shaft flooding. These occurred mainly during the sinking phase. There are many methods for eliminating water hazards and dewatering flooded shafts to put them back into operation. This paper presents two cases of highly problematic flooding in shaft sunk through highly waterlogged layers using the freezing method, both leading to considerable construction delays. The first case involved the removal of increased water inflow into the R-XI shaft

#### **Chapter 6**

## Polish Experiences in Handling Water Hazards during Mine Shaft Sinking

*Piotr Czaja, Paweł Kamiński and Artur Dyczko*

#### **Abstract**

The geological structure of most Polish mining regions is rich in groundwater, making shaft sinking difficult. In recent years, more than a dozen shafts, some almost 700 m deep, have been sunk in Poland using various methods of water hazard elimination. The vast majority of shafts that pass through aquifer formations have been sunk using artificial rock freezing, waterproof tubing, and concrete lining. Generally, this system has proven to be very effective. However, there have been cases of complications during sinking, including occasional flooding. This paper presents two cases of highly problematic flooding in shaft sunk using the freezing method, both leading to considerable construction delays and a significant increase in shaft sinking costs. The first case involved water inflow into the bottom section of the R-XI shaft at KGHM with rocks near the melting point of ice. In the other case, problems occurred passing through an Albian layer in the S. 1.3 shaft sunk for the Lubelski Węgiel Bogdanka S.A. mining corporation, where the freezing process was carried out while it was necessary to heat the rocks in the upper part of the shaft to protect the final lining from damage.

**Keywords:** mining shaft, water hazard, grouting, dewatering

#### **1. Introduction**

Exposing deep-seated mineral deposits requires the construction of new shafts. In Poland, where usable minerals are usually covered by thick layers of heavily waterlogged overburden, the construction of new shafts poses extraordinary difficulties. New shafts continue to be designed and constructed in quite challenging hydrogeological conditions in Poland, as well as in other countries worldwide. Hence, it would be fruitful to look at some Polish experiences in coping with this extremely difficult hydrogeology while mining deposits of both hard-coal and nonferrous metal ores. A range of detailed examples of how to eliminate such water hazards has been provided elsewhere [1–3]. Over the last three decades, Poland has seen at least several cases involving shaft flooding. These occurred mainly during the sinking phase. There are many methods for eliminating water hazards and dewatering flooded shafts to put them back into operation. This paper presents two cases of highly problematic flooding in shaft sunk through highly waterlogged layers using the freezing method, both leading to considerable construction delays. The first case involved the removal of increased water inflow into the R-XI shaft

**96**

*Mining Techniques - Past, Present and Future*

[1] Kostrz J, Olszewski J, Czaja P, Deja J, Witosiński J. Zastosowanie betonów odpornych na silną agresję siarczanową

podziemnym. Budownictwo Górnicze i Tunelowe. 2000;**3**:4-11. ISSN: 1234-5342

[2] Olszewski J, Czaja P, Bulenda P, Kamiński P. Pogłębianie oraz wydłużanie górniczych wyciągów szybowych—szybu Leon IV w kopalni KWK ROW—Ruch Rydułtowy. Przegląd

Górniczy. 2018;**74**(8):7-16. ISSN:

[3] Czaja P, Kamiński P. Wybrane zagadnienia techniki i technologii głębienia szybów. Kraków: Szkoła Eksploatacji Podziemnej. 2016. p. 161.

[4] Kicki J, Sobczyk EJ, Kamiński P. Vertical and decline shaft sinking— Good practices in technique and technology. In: International Mining Forum 2015; 23-27 February 2015; Cracow, Poland. Leiden: CRC Press/Balkema; 2015. p. 197. ISBN:

[5] Płachno M. Mathematical model of transverse vibration of a high-capacity mining skip due misalignment of the guiding tracks in the hoisting shaft. Archives of Mining Sciences (Archiwum Górnictwa). Wyd. Polska Akademia Nauk, Komitet Górnictwa, Kraków. 2018;**63**(1):5. DOI: 10.24425/118882

ISBN: 978-83-927920-9-3

978-1-138-02820-3

0033-216X

**References**

i magnezową w budownictwie

at KGHM. In the other case, problems occurred due to a shaft passing through an Albian layer in the S. 1.3 shaft sunk for the Lubelski Węgiel Bogdanka S.A. mining corporation. Although completely different from each other, these cases provide useful guidance and a serious warning against hasty shaft design or a careless approach to constructing shafts [4–6]. Considered completely safe for shaft construction, the technological solutions presented here should be of interest to experts in water-related mining issues.

#### **2. Diversion of increased water inflow into the R-XI shaft during sinking**

Waterlogged overburden formations as deep as 700 m below the ground have made it necessary for Polish mining corporations to use the freezing method to construct all copper mine shafts and most hard-coal mine shafts. Hundreds of shafts have been successfully sunk in Poland using this technology. However, when it seemed that the engineers had virtually eliminated freezing pipe leaks in the boreholes, a major problem that had caused brine leaks into frozen rock, water hazards emerged in completely unexpected and highly unlikely situations.

#### **2.1 Project specification and the effects of the water hazard**

The R-XI shaft was not the first structure of this type constructed by PeBeKa S.A. in the Polish Copper Basin area [7]. Hydrogeological surveys preceding the shaft work at depths of 431.0–630.0 m indicated no significant water hazards along this section. The projected water inflows into the shaft face below the 431.0 m level are shown in **Table 1**. The R-XI shaft was designed to serve as a ventilation shaft and has the following parameters [5]:


At the time, this shaft had the greatest rock freezing depth at 635 m. PeBeKa Lubin applied many innovative rock freezing solutions. One of them was selective freezing using two types of freezing holes: short holes with a depth of 395 m and


**99**

**Figure 1.**

*(d = 3.5," L = 576 m); 8, final concrete lining.*

*Polish Experiences in Handling Water Hazards during Mine Shaft Sinking*

long holes with a depth of 635 m. This method made it possible to achieve a frozen mantle that was thickest in its lower portion, where the water pressure was found to

Because a gallery had already been excavated near the shaft at a depth of 1212.7 m, the design included a simplified drainage system for the shaft face below the freezing zone. This was achieved through a dewatering borehole drilled in the shaft axis vertically upwards from a level of 1212.7 m. This made it possible to dispense with the construction of an expensive cascade drainage system and significantly facilitated shaft sinking at depths of 635–1212.7 m. At the 503.6–632.4 m shaft section, the design included a combined panel and concrete lining, i.e., a topdown panel lining and a concrete, monolithic, bottom-up lining using panel forms. The concrete lining was laid on a 2.6-m-thick base ring beam set between depths of

*The last phases of shaft sinking in the frozen rock area. (a) Installation of foundation for the final shaft concrete lining. (b) Section of final shaft concrete lining with drainage. Explanations: 1, three-deck shaft working platform; 2, cast-iron shaft lining; 3, preliminary pre-cast segmental shaft lining; 4, shaft lining foundation; 5, boreholes in the drainage system; 6, sliding formwork H = 3.75 m; 7, dewatering borehole TS-1* 

*DOI: http://dx.doi.org/10.5772/intechopen.93146*

be the highest.

632.4 and 635.0 m (**Figure 1**).

#### **Table 1.**

*Predicted water inflows into the shaft [7].*

*Polish Experiences in Handling Water Hazards during Mine Shaft Sinking DOI: http://dx.doi.org/10.5772/intechopen.93146*

long holes with a depth of 635 m. This method made it possible to achieve a frozen mantle that was thickest in its lower portion, where the water pressure was found to be the highest.

Because a gallery had already been excavated near the shaft at a depth of 1212.7 m, the design included a simplified drainage system for the shaft face below the freezing zone. This was achieved through a dewatering borehole drilled in the shaft axis vertically upwards from a level of 1212.7 m. This made it possible to dispense with the construction of an expensive cascade drainage system and significantly facilitated shaft sinking at depths of 635–1212.7 m. At the 503.6–632.4 m shaft section, the design included a combined panel and concrete lining, i.e., a topdown panel lining and a concrete, monolithic, bottom-up lining using panel forms. The concrete lining was laid on a 2.6-m-thick base ring beam set between depths of 632.4 and 635.0 m (**Figure 1**).

#### **Figure 1.**

*Mining Techniques - Past, Present and Future*

in water-related mining issues.

has the following parameters [5]:

• Lining diameter—7.5 m

• Aquifer thill depth—630 m

• Freezing depth—635 m

*Predicted water inflows into the shaft [7].*

• Total depth—1250 m

at KGHM. In the other case, problems occurred due to a shaft passing through an Albian layer in the S. 1.3 shaft sunk for the Lubelski Węgiel Bogdanka S.A. mining corporation. Although completely different from each other, these cases provide useful guidance and a serious warning against hasty shaft design or a careless approach to constructing shafts [4–6]. Considered completely safe for shaft construction, the technological solutions presented here should be of interest to experts

**2. Diversion of increased water inflow into the R-XI shaft during sinking**

Waterlogged overburden formations as deep as 700 m below the ground have made it necessary for Polish mining corporations to use the freezing method to construct all copper mine shafts and most hard-coal mine shafts. Hundreds of shafts have been successfully sunk in Poland using this technology. However, when it seemed that the engineers had virtually eliminated freezing pipe leaks in the boreholes, a major problem that had caused brine leaks into frozen rock, water hazards emerged in completely unexpected and highly unlikely situations.

The R-XI shaft was not the first structure of this type constructed by PeBeKa S.A. in the Polish Copper Basin area [7]. Hydrogeological surveys preceding the shaft work at depths of 431.0–630.0 m indicated no significant water hazards along this section. The projected water inflows into the shaft face below the 431.0 m level are shown in **Table 1**. The R-XI shaft was designed to serve as a ventilation shaft and

At the time, this shaft had the greatest rock freezing depth at 635 m. PeBeKa Lubin applied many innovative rock freezing solutions. One of them was selective freezing using two types of freezing holes: short holes with a depth of 395 m and

431.0–460.0 0.042–0.070 0.056 460.0–470.0 0.042–0.070 0.056 470.0–500.0 0.061–0.330 0.160 500.0–565.0 0.205–0.490 0.334 565.0–630.0 0.334–0.550 0.425

**/min]**

**Minimum to maximum Average**

**Depth interval [m] Water inflow [m3**

**2.1 Project specification and the effects of the water hazard**

**98**

**Table 1.**

*The last phases of shaft sinking in the frozen rock area. (a) Installation of foundation for the final shaft concrete lining. (b) Section of final shaft concrete lining with drainage. Explanations: 1, three-deck shaft working platform; 2, cast-iron shaft lining; 3, preliminary pre-cast segmental shaft lining; 4, shaft lining foundation; 5, boreholes in the drainage system; 6, sliding formwork H = 3.75 m; 7, dewatering borehole TS-1 (d = 3.5," L = 576 m); 8, final concrete lining.*

According to the records [6, 7], the in situ rock temperature at a depth of 632 m was about 32°C. So, it was reasonable to expect that the end portion of the frozen mantle would also be exposed to increased heat from below. When the shaft face reached a depth of 632 m without any difficulties, it seemed that the most challenging section had been sunk as designed and on schedule. Yet, nature retained its unpredictability.

After the two last reinforced concrete panel rings had been completed, with excess material excavated to make a curb gap 4 (**Figure 1**) at a depth of 632 m, a small water leak, estimated at about 3–5 L/min, was noticed at the shaft bottom at the thill sidewall interface. The water was clean, very cold, and slightly salty. For a shaft sunk using the freezing method, in which the freezing core usually has a temperature below −15°C, this was unusual and perplexing. Since the freezing pipes had reached a depth of 635 m, no liquid water should have occurred at a depth of 632 m. However, this phenomenon could be partly explained by the water's salinity. Unfortunately, the electrical conductivity of this water has not been documented. In these circumstances, the TS-1 dewatering borehole work was intensified. Also, work commenced on the final concrete lining 8 (**Figure 1**)—constructed from the bottom up—equipped with a drainage system [7].

It was found that even though all the freezing safety requirements had been observed, the ice mantle along this section was not completely watertight and did not fully prevent water inflow into the shaft face. The movement of slightly saline water at a temperature above zero (tw > 0°C) caused the frozen mantle to be soaked from below and consistently thawed, with water inflows effectively increasing day by day. The situation was becoming dangerous, as no shaft pipe drainage had been planned down to this depth. This meant that the shaft had no pipelines through which the water could be pumped up to the surface. The further section of the shaft was designed to allow drainage via the TS-1 dewatering borehole drilled from a level of 1212.7 m (**Figure 2**).

The increasing inflow of water was diverted to the surface using only buckets. After about 2 weeks of shaft work involving the construction of a concrete curb at a depth of 635 m and the construction of an 18 m final concrete lining, water inflow into the shaft had increased to about 700 L/min. In this situation, it was impossible to continue any work in the shaft other than intensive dewatering using of buckets. Ultimately, this measure did not save the shaft from partial flooding. The water table in the shaft stabilized at a depth of 533.0 m, which means that the water column was 102 m (see **Figure 2**).

Due to the prolonged length of the 564 m TS-1 dewatering borehole and the water level reaching 533 m (**Figure 2a**), the decision was made to use a highperformance RITZ submersible pump (HDM 6723/11DPF). Installed 4 weeks later, with a capacity of 15 m3 /min, the submersible pump succeeded in quickly dewatering the flooded shaft section (**Figure 2b**). Also, after 2 months of further work, the water inflow into the shaft was found to have reached 2.5 m3 /min. The dewatering borehole TS-1 (**Figure 2**) was successfully completed almost at the same time the shaft was dewatered using the submersible pump. After 6 weeks of intensive and highly precise drilling work, the borehole reached the shaft bottom, located only 0.5 m from the shaft axis. By this point, the water inflow had increased to 3.0 m3 / min. Since the water inflow was expected to increase further, the decision was made to drill a second dewatering borehole—TS-2 (**Figure 2c**). Due to the considerable water hazard associated with a water inflow of 3.0 m3 /min, it was also decided that the section with a waterproof tubing lining be extended to the 650 m level. In addition, the decision was made to comprehensively grout the entire area affected by the substantial water inflow.

**101**

**Figure 2.**

*dewatering borehole.*

*Polish Experiences in Handling Water Hazards during Mine Shaft Sinking*

*DOI: http://dx.doi.org/10.5772/intechopen.93146*

**2.2 Removing the causes of the water inflow**

The substantial water inflow forced the shaft construction company to both further redesign the shaft lining and adjust the sinking technology along the 635–650-m-deep section. Apart from the costly dewatering, one of the direct effects of the partial shaft flooding was the need to redesign the lining in the flooding area (**Figure 2**). The concrete panel lining was replaced by a tubing panel lining, with a concrete tube set between them (**Figure 2c**) [7]. Due to this replacement, it was additionally necessary to:

*The phases of dewatering the sunken shaft section. (a) Shaft flooding. (b) Dewatering of the shaft using a "RITZ" submersible pump, (c) replacing the concrete lining along the 574.3–632.9 m section with a cast-iron tubing lining. Explanations: 1–8, see Figure 1; 9, RITZ submersible pump; 10, cast-iron tubing lining; 11, TS-2* 

a.Demolish the completed 18 m section of the concrete lining above the curb, at a

b.Partially demolish the curb at a depth of 635 m and mount a steel ring beam on

depth of 635 m, without damaging the preliminary panel lining.

the curb's foundations to lay the first tubing ring.

*Polish Experiences in Handling Water Hazards during Mine Shaft Sinking DOI: http://dx.doi.org/10.5772/intechopen.93146*

#### **Figure 2.**

*Mining Techniques - Past, Present and Future*

unpredictability.

system [7].

of 1212.7 m (**Figure 2**).

with a capacity of 15 m3

substantial water inflow.

column was 102 m (see **Figure 2**).

According to the records [6, 7], the in situ rock temperature at a depth of 632 m was about 32°C. So, it was reasonable to expect that the end portion of the frozen mantle would also be exposed to increased heat from below. When the shaft face reached a depth of 632 m without any difficulties, it seemed that the most challenging section had been sunk as designed and on schedule. Yet, nature retained its

After the two last reinforced concrete panel rings had been completed, with excess material excavated to make a curb gap 4 (**Figure 1**) at a depth of 632 m, a small water leak, estimated at about 3–5 L/min, was noticed at the shaft bottom at the thill sidewall interface. The water was clean, very cold, and slightly salty. For a shaft sunk using the freezing method, in which the freezing core usually has a temperature below −15°C, this was unusual and perplexing. Since the freezing pipes had reached a depth of 635 m, no liquid water should have occurred at a depth of 632 m. However, this phenomenon could be partly explained by the water's salinity. Unfortunately, the electrical conductivity of this water has not been documented. In these circumstances, the TS-1 dewatering borehole work was intensified. Also, work commenced on the final concrete lining 8 (**Figure 1**)—constructed from the bottom up—equipped with a drainage

It was found that even though all the freezing safety requirements had been observed, the ice mantle along this section was not completely watertight and did not fully prevent water inflow into the shaft face. The movement of slightly saline water at a temperature above zero (tw > 0°C) caused the frozen mantle to be soaked from below and consistently thawed, with water inflows effectively increasing day by day. The situation was becoming dangerous, as no shaft pipe drainage had been planned down to this depth. This meant that the shaft had no pipelines through which the water could be pumped up to the surface. The further section of the shaft was designed to allow drainage via the TS-1 dewatering borehole drilled from a level

The increasing inflow of water was diverted to the surface using only buckets. After about 2 weeks of shaft work involving the construction of a concrete curb at a depth of 635 m and the construction of an 18 m final concrete lining, water inflow into the shaft had increased to about 700 L/min. In this situation, it was impossible to continue any work in the shaft other than intensive dewatering using of buckets. Ultimately, this measure did not save the shaft from partial flooding. The water table in the shaft stabilized at a depth of 533.0 m, which means that the water

Due to the prolonged length of the 564 m TS-1 dewatering borehole and the water level reaching 533 m (**Figure 2a**), the decision was made to use a highperformance RITZ submersible pump (HDM 6723/11DPF). Installed 4 weeks later,

ing the flooded shaft section (**Figure 2b**). Also, after 2 months of further work, the

borehole TS-1 (**Figure 2**) was successfully completed almost at the same time the shaft was dewatered using the submersible pump. After 6 weeks of intensive and highly precise drilling work, the borehole reached the shaft bottom, located only 0.5 m from the shaft axis. By this point, the water inflow had increased to 3.0 m3

min. Since the water inflow was expected to increase further, the decision was made to drill a second dewatering borehole—TS-2 (**Figure 2c**). Due to the considerable

the section with a waterproof tubing lining be extended to the 650 m level. In addition, the decision was made to comprehensively grout the entire area affected by the

water inflow into the shaft was found to have reached 2.5 m3

water hazard associated with a water inflow of 3.0 m3

/min, the submersible pump succeeded in quickly dewater-

/min. The dewatering

/min, it was also decided that

/

**100**

*The phases of dewatering the sunken shaft section. (a) Shaft flooding. (b) Dewatering of the shaft using a "RITZ" submersible pump, (c) replacing the concrete lining along the 574.3–632.9 m section with a cast-iron tubing lining. Explanations: 1–8, see Figure 1; 9, RITZ submersible pump; 10, cast-iron tubing lining; 11, TS-2 dewatering borehole.*

#### **2.2 Removing the causes of the water inflow**

The substantial water inflow forced the shaft construction company to both further redesign the shaft lining and adjust the sinking technology along the 635–650-m-deep section. Apart from the costly dewatering, one of the direct effects of the partial shaft flooding was the need to redesign the lining in the flooding area (**Figure 2**). The concrete panel lining was replaced by a tubing panel lining, with a concrete tube set between them (**Figure 2c**) [7]. Due to this replacement, it was additionally necessary to:


In the first phase, the rock behind the lining was grouted using multiple techniques. In the first phase, 3-m-long holes were drilled in rings 309 and 310 through cement plugs in the tubing lining. A total of 26 t of cement grout were injected behind the lining through these holes to separate the upper water horizons from the problem area of the shaft.

In the second phase, the cement grout was injected behind the lining along the 617.9–635.0 m section, using 2-m-long horizontal holes drilled through the concrete plugs, 10-m-long horizontal holes drilled through the cement plugs, and 15-m-long inclined holes drilled at an angle of 40° through the concrete plugs. Due to the very substantial water inflow from this area, "Ekopur HW" quickset two-component

#### **Figure 3.**

*Grouting process and shaft sinking along the 635–650 m section. (a) Grouting along the 598.4–635.0 m section, phase 1 and 2; (b) shaft sinking along the 635–650 m section, (c) grouting along the 635–650 m section, phase 3. Explanations: 1, cementation of the rock behind the lining (insulating layer) in N-130a tubing rings 309 and 310; 2, grouting of the rock behind the tubing lining through concrete plug holes in the tubing; 3, grouting of the rock and tubing lining through "cementation" holes in tubings (2-m-long horizontal holes); 4, grouting of the rock and tubing lining through "cementation" holes in tubings (10-m-long horizontal holes); 5, grouting of the rock and tubing lining through concrete plug holes (15-m-long inclined holes); 6, grouting of the rock and tubing lining through "cementation" holes in tubings (2.0-m-long inclined holes), 7, grouting of the rock and tubing lining through "cementation" holes in tubings (10.0-m-long inclined holes).*

**103**

*Polish Experiences in Handling Water Hazards during Mine Shaft Sinking*

**Phase Grouting materials used [Mg]**

polyurethane adhesive was used in addition to the cement grout. The grouting work is illustrated in **Figure 3** [7]. Once the 635–650 m section of the shaft had been sunk, a curb was made in the tubing lining, comprising 130a tubings (9 rings) installed from the top down at a depth of 650 m, and a shaft face dewatering system

Phase 1 45.5 24.5 70.0 Phase 2 33.3 14.8 48.1 Phase 3 16.4 0.32 16.7 Phase 4 342.7 32.4 375.1 **Total 438.0 72.1 510.1**

**Cement EKOPUR HW polyurethane Total**

In the third phase, cement grout was injected behind the lining along the 635–635.0 m section (**Table 2**). Then, as part of the fourth grouting phase, the entire 574–500 m section of the tubing lining was sealed. It took a total of more than

**3. Eliminating water hazards associated with the S-1.3 shaft sinking** 

The hydrogeology of the Lublin Coal Basin is highly complex, and mining in this area is challenging. At 710 m from the surface, the coal measures are covered by heavily waterlogged Jurassic, Cretaceous, and Quaternary formations. This has considerable implications for mine shaft sinking. A simplified geological profile is

All the shafts in this basin had to be sunk using the freezing method, at least along the 0–180 m section [4, 6, 8]. The first shafts for the Bogdanka Mine were given the numbers S-1.1, S-1.2, and S-1.3. After the S-1.1 shaft had been sunk to a depth of 960 m, a disastrous water leakage occurred from the connector pipes left in the lining, which caused extensive flooding. Consequently, it was necessary to fill in and abandon that shaft. Drawing on the S-1.1 experience, the S-1.2 shaft was sunk to the target depth of 995 m without any major difficulties. Although the flooding of the S-1.1 shaft had also caused partial flooding of the S-1.2 shaft through the galleries already sunk to a depth of 960 m, the dewatering proved to be fairly easy. The sinking of the S-1.3 shaft might be the most interesting and perhaps the only such case in the global history of shaft construction, as it ultimately required simultaneous rock freezing in the lower section and rock heating in the

The experience gained sinking the S-1.1 and S-1.2 shafts indicated that it was possible to use a different technology, more based on the traditional sinking method, which is much less costly. A decision was made to freeze the rocks along the 0–180 m section before constructing the first section, as it passed through the Quaternary strata and the highly waterlogged layers of Cretaceous formations, with a maximum

upper section. Below is a detailed discussion of how this was done.

was installed using boreholes TS-1 and TS-2 (**Figure 2**).

**project in the Lublin Coal Basin**

presented in **Figure 4**.

**Table 2.**

**3.1 The S-1.3 shaft sinking**

500 t of materials (**Table 2**) to complete the grouting process.

*DOI: http://dx.doi.org/10.5772/intechopen.93146*

*Grouting materials used to prevent water inflow [7].*


*Polish Experiences in Handling Water Hazards during Mine Shaft Sinking DOI: http://dx.doi.org/10.5772/intechopen.93146*

#### **Table 2.**

*Mining Techniques - Past, Present and Future*

ing a curb at a depth of 650 m.

problem area of the shaft.

tubing column at a depth of 500.47 m.

c.Construct the lining of 39 N-130a tubing rings from the bottom up, to a depth of 574.3 m, without damaging the preliminary panel lining, the initial step being to lay the first ring on the steel ring beam in the curb at a depth of 635 m.

d.Complete the grouting work above the curb, at a depth of 635 m, so that the

e. Sink the shaft along the 635–650 m tubing-lined section, including construct-

f. Construct the lining of N-120 cast-iron tubing rings from the 574.3 m level upwards to the point of connection between the picotage gap and the upper

In the first phase, the rock behind the lining was grouted using multiple techniques. In the first phase, 3-m-long holes were drilled in rings 309 and 310 through cement plugs in the tubing lining. A total of 26 t of cement grout were injected behind the lining through these holes to separate the upper water horizons from the

In the second phase, the cement grout was injected behind the lining along the 617.9–635.0 m section, using 2-m-long horizontal holes drilled through the concrete plugs, 10-m-long horizontal holes drilled through the cement plugs, and 15-m-long inclined holes drilled at an angle of 40° through the concrete plugs. Due to the very substantial water inflow from this area, "Ekopur HW" quickset two-component

*Grouting process and shaft sinking along the 635–650 m section. (a) Grouting along the 598.4–635.0 m section, phase 1 and 2; (b) shaft sinking along the 635–650 m section, (c) grouting along the 635–650 m section, phase 3. Explanations: 1, cementation of the rock behind the lining (insulating layer) in N-130a tubing rings 309 and 310; 2, grouting of the rock behind the tubing lining through concrete plug holes in the tubing; 3, grouting of the rock and tubing lining through "cementation" holes in tubings (2-m-long horizontal holes); 4, grouting of the rock and tubing lining through "cementation" holes in tubings (10-m-long horizontal holes); 5, grouting of the rock and tubing lining through concrete plug holes (15-m-long inclined holes); 6, grouting of the rock and tubing lining through "cementation" holes in tubings (2.0-m-long inclined holes), 7, grouting of the rock and tubing lining through "cementation" holes in* 

shaft could be safely sunk along the 635–650 m interval.

**102**

*tubings (10.0-m-long inclined holes).*

**Figure 3.**

*Grouting materials used to prevent water inflow [7].*

polyurethane adhesive was used in addition to the cement grout. The grouting work is illustrated in **Figure 3** [7]. Once the 635–650 m section of the shaft had been sunk, a curb was made in the tubing lining, comprising 130a tubings (9 rings) installed from the top down at a depth of 650 m, and a shaft face dewatering system was installed using boreholes TS-1 and TS-2 (**Figure 2**).

In the third phase, cement grout was injected behind the lining along the 635–635.0 m section (**Table 2**). Then, as part of the fourth grouting phase, the entire 574–500 m section of the tubing lining was sealed. It took a total of more than 500 t of materials (**Table 2**) to complete the grouting process.

#### **3. Eliminating water hazards associated with the S-1.3 shaft sinking project in the Lublin Coal Basin**

The hydrogeology of the Lublin Coal Basin is highly complex, and mining in this area is challenging. At 710 m from the surface, the coal measures are covered by heavily waterlogged Jurassic, Cretaceous, and Quaternary formations. This has considerable implications for mine shaft sinking. A simplified geological profile is presented in **Figure 4**.

All the shafts in this basin had to be sunk using the freezing method, at least along the 0–180 m section [4, 6, 8]. The first shafts for the Bogdanka Mine were given the numbers S-1.1, S-1.2, and S-1.3. After the S-1.1 shaft had been sunk to a depth of 960 m, a disastrous water leakage occurred from the connector pipes left in the lining, which caused extensive flooding. Consequently, it was necessary to fill in and abandon that shaft. Drawing on the S-1.1 experience, the S-1.2 shaft was sunk to the target depth of 995 m without any major difficulties. Although the flooding of the S-1.1 shaft had also caused partial flooding of the S-1.2 shaft through the galleries already sunk to a depth of 960 m, the dewatering proved to be fairly easy. The sinking of the S-1.3 shaft might be the most interesting and perhaps the only such case in the global history of shaft construction, as it ultimately required simultaneous rock freezing in the lower section and rock heating in the upper section. Below is a detailed discussion of how this was done.

#### **3.1 The S-1.3 shaft sinking**

The experience gained sinking the S-1.1 and S-1.2 shafts indicated that it was possible to use a different technology, more based on the traditional sinking method, which is much less costly. A decision was made to freeze the rocks along the 0–180 m section before constructing the first section, as it passed through the Quaternary strata and the highly waterlogged layers of Cretaceous formations, with a maximum

#### **Figure 4.**

*Diagram of the S-1.3 shaft sinking using both rock freezing and rock heating. 1, boreholes for rock heating along the 0–180 m section; 2, boreholes for deep freezing along the 0–570 m section; 3, concrete plug above the Albian layer; 4, drainage boreholes in the Albian layer; 5, working with a drilling chamber at a depth of 754 m; 6, frozen rock mantle; 7, final concrete panel lining; 8, cast-iron tubing lining along the Albian formation section.*

depth of 162 m (see **Figure 4**). Below the 180 m level, the plan was to sink the shaft to a depth of 570 m using conventional method, i.e., without rock freezing. This would substantially reduce costs. The biggest puzzle, and, as the construction company would see, the greatest challenge involved in this shaft sinking project was the thin (≈2.9 m) Albian layer (**Figure 4**), which was composed of sandy-lime quicksand with a water pressure of about 5.5 MPa. An assumption was made that a shaft working could pass through such a thin layer of waterlogged formation once the layer had been provided with borehole drainage system (**Figure 4**) drilled in the working at a depth of 754 m. With the drainage system in place, the pressure could be reduced, making it possible to petrify both the Albian formations and the Jurassic formations deposited underneath, all the way to the Carboniferous roof. This way, the shaft could be sunk conventionally down to the target depth of 1035.45 m.

Here, we should warn those who are enthusiastic about using grouting, regardless of the conditions. In this specific case, the company constructing the shaft failed to provide the mentioned formations with a drainage system. In effect, it became impossible to chemically petrify the Jurassic formations any further, and the only viable sinking option left was the freezing method. At this

**105**

*Polish Experiences in Handling Water Hazards during Mine Shaft Sinking*

Heat supplied through the boreholes along the 14 m diameter circle

point, the expected substantial savings stemming from the use of a different shaft sinking method were no longer viable. In addition, the construction company faced the problem of refreezing within the 0–570 m zone, where the final lining had already been laid. The Polish engineers involved in the project knew that the refreezing of rocks would produce a great pressure surge on the lining, effectively

*Projected and actual energy consumption in the process of rock freezing when sinking the S-1.3 shaft [3].*

**Parameter Design Actual** Freezing time, months 9.2 About 7 Amount of energy consumed to create the frozen mantle, MJ 71,310,951 45,638,000

Heat supplied through the air supply duct to the warm-air shaft, MJ 23,197,000 16,130,000 **Total energy consumed, MJ** 109,830,951 74,277,800

15,323,000 12,509,800

The engineers considered it necessary to drill 43 additional boreholes at a depth of 610 m. These had an unusual diameter of 308 mm and were drilled in an 18 m diameter circle [3, 8]. Also, an unprecedented decision was made to use sectional freezing—an approach which, although theoretically known and viable, had not been applied in shaft construction before. Thus, the boreholes were fitted with two freezing pipe columns and a column of downcomer tubes inside a 139.7 mm diameter column. They were properly sealed so that the brine could circulate only in the

Regrettably, this plan failed, too. The shrinkage stress in the steel due to the low temperature of the brine caused the outer column to leak, allowing water to enter the borehole. This complicated the whole process of section freezing, making it necessary to reconsider freezing along the entire depth of the shaft. As feared, the freezing caused damage to the lining along the 0–570 m section soon after commenced. At this point, the decision was made to apply a globally unprecedented solution, in which the lower section of the shaft was frozen, while the upper part of the shaft, along the 0–180 m section, was heated with warm water. To provide the inflow of warm water, the engineers used the boreholes drilled to freeze the first section of the shaft along a circle with a diameter of 14 m. The work diagram is

Ultimately, this unprecedented project proved a technological success. However, although the shaft was eventually sunk, the project can hardly be described as successful, given the completion period of almost 10 years and the substantial energy costs involved. The substantial costs of sinking the S-1.3 shaft are reflected in the amount of energy consumed in the process of rock freezing and heating. These parameters are shown in **Table 3**. It should be noted that the actual values were

This paper shows how changeable and unpredictable hydrogeology can lead to challenging and very costly problems in shaft sinking projects. In the case of the polish shafts, a water hazard that had not been accurately identified by hydrogeological surveys led to a number of adverse effects. These included the substantial amount of grouting materials used, the extended project completion period (it took

*DOI: http://dx.doi.org/10.5772/intechopen.93146*

lower parts of the boreholes, below 570 m.

much lower (by about 38%) than the design values.

destroying it [6].

(0–180 m), MJ

**Table 3.**

presented in **Figure 4**.

**4. Conclusions**

*Polish Experiences in Handling Water Hazards during Mine Shaft Sinking DOI: http://dx.doi.org/10.5772/intechopen.93146*


**Table 3.**

*Mining Techniques - Past, Present and Future*

depth of 162 m (see **Figure 4**). Below the 180 m level, the plan was to sink the shaft to a depth of 570 m using conventional method, i.e., without rock freezing. This would substantially reduce costs. The biggest puzzle, and, as the construction company would see, the greatest challenge involved in this shaft sinking project was the thin (≈2.9 m) Albian layer (**Figure 4**), which was composed of sandy-lime quicksand with a water pressure of about 5.5 MPa. An assumption was made that a shaft working could pass through such a thin layer of waterlogged formation once the layer had been provided with borehole drainage system (**Figure 4**) drilled in the working at a depth of 754 m. With the drainage system in place, the pressure could be reduced, making it possible to petrify both the Albian formations and the Jurassic formations deposited underneath, all the way to the Carboniferous roof. This way, the shaft could be sunk conventionally down to the target depth of 1035.45 m. Here, we should warn those who are enthusiastic about using grouting, regardless of the conditions. In this specific case, the company constructing the shaft failed to provide the mentioned formations with a drainage system. In effect, it became impossible to chemically petrify the Jurassic formations any further, and the only viable sinking option left was the freezing method. At this

*Diagram of the S-1.3 shaft sinking using both rock freezing and rock heating. 1, boreholes for rock heating along the 0–180 m section; 2, boreholes for deep freezing along the 0–570 m section; 3, concrete plug above the Albian layer; 4, drainage boreholes in the Albian layer; 5, working with a drilling chamber at a depth of 754 m; 6, frozen rock mantle; 7, final concrete panel lining; 8, cast-iron tubing lining along the Albian formation section.*

**104**

**Figure 4.**

*Projected and actual energy consumption in the process of rock freezing when sinking the S-1.3 shaft [3].*

point, the expected substantial savings stemming from the use of a different shaft sinking method were no longer viable. In addition, the construction company faced the problem of refreezing within the 0–570 m zone, where the final lining had already been laid. The Polish engineers involved in the project knew that the refreezing of rocks would produce a great pressure surge on the lining, effectively destroying it [6].

The engineers considered it necessary to drill 43 additional boreholes at a depth of 610 m. These had an unusual diameter of 308 mm and were drilled in an 18 m diameter circle [3, 8]. Also, an unprecedented decision was made to use sectional freezing—an approach which, although theoretically known and viable, had not been applied in shaft construction before. Thus, the boreholes were fitted with two freezing pipe columns and a column of downcomer tubes inside a 139.7 mm diameter column. They were properly sealed so that the brine could circulate only in the lower parts of the boreholes, below 570 m.

Regrettably, this plan failed, too. The shrinkage stress in the steel due to the low temperature of the brine caused the outer column to leak, allowing water to enter the borehole. This complicated the whole process of section freezing, making it necessary to reconsider freezing along the entire depth of the shaft. As feared, the freezing caused damage to the lining along the 0–570 m section soon after commenced. At this point, the decision was made to apply a globally unprecedented solution, in which the lower section of the shaft was frozen, while the upper part of the shaft, along the 0–180 m section, was heated with warm water. To provide the inflow of warm water, the engineers used the boreholes drilled to freeze the first section of the shaft along a circle with a diameter of 14 m. The work diagram is presented in **Figure 4**.

Ultimately, this unprecedented project proved a technological success. However, although the shaft was eventually sunk, the project can hardly be described as successful, given the completion period of almost 10 years and the substantial energy costs involved. The substantial costs of sinking the S-1.3 shaft are reflected in the amount of energy consumed in the process of rock freezing and heating. These parameters are shown in **Table 3**. It should be noted that the actual values were much lower (by about 38%) than the design values.

#### **4. Conclusions**

This paper shows how changeable and unpredictable hydrogeology can lead to challenging and very costly problems in shaft sinking projects. In the case of the polish shafts, a water hazard that had not been accurately identified by hydrogeological surveys led to a number of adverse effects. These included the substantial amount of grouting materials used, the extended project completion period (it took almost an additional year to finish the project), and the need to replace the concrete panel lining with tubing lining along a 150-m-long section of the shaft.

This paper presents case histories that should serve as the ultimate warning against underestimating the projected inflow of water into a shaft during its sinking. A number of shaft construction projects recently implemented in Poland further illustrate this point. Since water inflow projections proved inaccurate, it is necessary to improve the accuracy of hydrogeological surveys in the areas where mining is planned. Poland has extensive and highly informative experience in successfully dealing with water hazards related to shaft construction.

#### **Author details**

Piotr Czaja1 , Paweł Kamiński1 \* and Artur Dyczko2

1 AGH University of Science and Technology, Kraków, Poland

2 Mineral and Energy Economy Research Institute, Polish Academy of Science, Kraków, Poland

\*Address all correspondence to: pkamin@agh.edu.pl

© 2020 The Author(s). Licensee IntechOpen. This chapter is distributed under the terms of the Creative Commons Attribution License (http://creativecommons.org/licenses/ by/3.0), which permits unrestricted use, distribution, and reproduction in any medium, provided the original work is properly cited.

**107**

pp. 32-38

*Polish Experiences in Handling Water Hazards during Mine Shaft Sinking*

[7] Kosmalski M, Kulicki J, Stróżyński M. The elimination of increased water inflows into the R-XI shaft in the Zakłady Górnicze Rudna mine. Geoinżynieria i Tunelowanie.

[8] Kicki J, Dyczko A. 30 years of LW "Bogdanka" S.A. In: The History and the

Future. Krakow: Bogdanka;2012

2005;**1**:46-54

*DOI: http://dx.doi.org/10.5772/intechopen.93146*

[1] Ciania Z, Lekan W, Wójcik J. Planned

[2] Czaja P, Kohutek Z, Wichur A. The water-related problems and difficulties encountered when exposing deposits – Coping methods. In: The Hydrogeology of Polish Mine Deposits and the Water-Related Problems in Mining. Kraków: Uczelniane Wydawnictwa Naukowo-Dydaktyczne AGH; 2003. pp. 146-200

[3] Kohutek Z, Wichur A, Wilk Z. Water-related issues when exposing deposits. In: Wilk Z, editor. The Hydrogeology of Polish Mine Deposits and the Water-Related Problems in Mining. Kraków: Uczelniane

Wydawnictwa Naukowo-Dydaktyczne

[4] Krywult J, Wichur A. A simplified method for designing preliminary linings for shafts sunk using rock freezing. Przegląd Górniczy; 1993. p. 5

[5] Stachowiak-Maciejowska K, Rożek R. The R-XI shaft is the 29th shaft of this Polish copper mining company, sunk by PeBeKa S.A. Nowoczesne Budownictwo Inżynieryjne. 2005;**09**:35-36. Available from: http://www.nbi.com.pl/assets/ NBI-pdf/2005/2/pdf/9\_pebeka.pdf

[6] Wichur A, Czaja P, Poprawski W. Guidelines for designing the frozen mantle thickness. OBRBG Budokop w Mysłowicach, Konferencja naukowo-techniczna Budownictwo górnicze i podziemne w nowych warunkach gospodarowania, Materiały konferencyjne i referaty problemowe, Kokotek k/Lublińca 16-17. IX. 1991.

AGH; 2003. pp. 424-435

and interventional cementation treatments in shafts sunk within the Lublin Coal Basin. In: Proc Symp: Experiences in the Use of the Grouting Technology in Rock Sealing, Reinforcing and Petrification in Underground Construction and Workings. Katowice:

**References**

SITG; 1988

*Polish Experiences in Handling Water Hazards during Mine Shaft Sinking DOI: http://dx.doi.org/10.5772/intechopen.93146*

#### **References**

*Mining Techniques - Past, Present and Future*

**106**

**Author details**

Kraków, Poland

, Paweł Kamiński1

provided the original work is properly cited.

\* and Artur Dyczko2

almost an additional year to finish the project), and the need to replace the concrete

This paper presents case histories that should serve as the ultimate warning against underestimating the projected inflow of water into a shaft during its sinking. A number of shaft construction projects recently implemented in Poland further illustrate this point. Since water inflow projections proved inaccurate, it is necessary to improve the accuracy of hydrogeological surveys in the areas where mining is planned. Poland has extensive and highly informative experience in

panel lining with tubing lining along a 150-m-long section of the shaft.

successfully dealing with water hazards related to shaft construction.

2 Mineral and Energy Economy Research Institute, Polish Academy of Science,

© 2020 The Author(s). Licensee IntechOpen. This chapter is distributed under the terms of the Creative Commons Attribution License (http://creativecommons.org/licenses/ by/3.0), which permits unrestricted use, distribution, and reproduction in any medium,

1 AGH University of Science and Technology, Kraków, Poland

\*Address all correspondence to: pkamin@agh.edu.pl

Piotr Czaja1

[1] Ciania Z, Lekan W, Wójcik J. Planned and interventional cementation treatments in shafts sunk within the Lublin Coal Basin. In: Proc Symp: Experiences in the Use of the Grouting Technology in Rock Sealing, Reinforcing and Petrification in Underground Construction and Workings. Katowice: SITG; 1988

[2] Czaja P, Kohutek Z, Wichur A. The water-related problems and difficulties encountered when exposing deposits – Coping methods. In: The Hydrogeology of Polish Mine Deposits and the Water-Related Problems in Mining. Kraków: Uczelniane Wydawnictwa Naukowo-Dydaktyczne AGH; 2003. pp. 146-200

[3] Kohutek Z, Wichur A, Wilk Z. Water-related issues when exposing deposits. In: Wilk Z, editor. The Hydrogeology of Polish Mine Deposits and the Water-Related Problems in Mining. Kraków: Uczelniane Wydawnictwa Naukowo-Dydaktyczne AGH; 2003. pp. 424-435

[4] Krywult J, Wichur A. A simplified method for designing preliminary linings for shafts sunk using rock freezing. Przegląd Górniczy; 1993. p. 5

[5] Stachowiak-Maciejowska K, Rożek R. The R-XI shaft is the 29th shaft of this Polish copper mining company, sunk by PeBeKa S.A. Nowoczesne Budownictwo Inżynieryjne. 2005;**09**:35-36. Available from: http://www.nbi.com.pl/assets/ NBI-pdf/2005/2/pdf/9\_pebeka.pdf

[6] Wichur A, Czaja P, Poprawski W. Guidelines for designing the frozen mantle thickness. OBRBG Budokop w Mysłowicach, Konferencja naukowo-techniczna Budownictwo górnicze i podziemne w nowych warunkach gospodarowania, Materiały konferencyjne i referaty problemowe, Kokotek k/Lublińca 16-17. IX. 1991. pp. 32-38

[7] Kosmalski M, Kulicki J, Stróżyński M. The elimination of increased water inflows into the R-XI shaft in the Zakłady Górnicze Rudna mine. Geoinżynieria i Tunelowanie. 2005;**1**:46-54

[8] Kicki J, Dyczko A. 30 years of LW "Bogdanka" S.A. In: The History and the Future. Krakow: Bogdanka;2012

**Chapter 7**

Anchorage Pile Strengthening of

Falling Stone Blocks by Mixture of

Melted Waste Plastics/Asphalt and

The Rock Fallings, Shale Slopes Stability, and Stability Risk Assessment in Şırnak open pit asphaltite mining should be searched in detail and improved in several coal mining sites in Şırnak Province, reaching over 120 m height with 60–65 degree shale slopes, developing major landslide in the open pit Şırnak open pit coal mining. The rock fallings endangered the mining safety in recent years. This research provided stability patterns and cementing method strengthening cracks. The stages of experimentation explored the geo-technical characteristics and geoological formation. Fort his aim, four different open pit mining areas with similar geotechnical conditions, two main strengthening methods, and patterns of researches were developed. Firstly, data on landslides and rock dynamics over explosions were followed, and secondly, as happened commonly in the past, the same geological, geomorphologic, hydrological, climatic conditions were taken. Anchorage pile strengthenıng of slopes and cementing falling stone blocks were performed by mixtures of melted waste plastics/asphalt and fly ash for stability of higher slopes over 120 m height and over 65 degree in asphaltite mining site in Silopi and Avgamasya open pit No.1 mining site in Şırnak were carried out. On the other hand, due to that creep style rock falling from top of slopes, those melted polymer cementing of anchorage bolting and cracks, to eliminate those falling failure types and features, will be advantageous. The unconditional expectations related to this study was also defined for this region, such as the influence of the ground water, rock cracks and slope design, and explosion exchange dynamics leading to landslide. GEO5 software and

Shale Slopes and Cementing

Fly Ash for Slope Stability in

in Avgamasya, Şırnak

manual stability analysis showed high risk area for plotting.

geotechnical stability, slope stability

**109**

**Keywords:** mining pit, Şırnak asphaltite, active potential landslide, mining,

*Yildırım İsmail Tosun*

**Abstract**

Asphaltite Open Pit Mining Site

### **Chapter 7**

Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture of Melted Waste Plastics/Asphalt and Fly Ash for Slope Stability in Asphaltite Open Pit Mining Site in Avgamasya, Şırnak

*Yildırım İsmail Tosun*

#### **Abstract**

The Rock Fallings, Shale Slopes Stability, and Stability Risk Assessment in Şırnak open pit asphaltite mining should be searched in detail and improved in several coal mining sites in Şırnak Province, reaching over 120 m height with 60–65 degree shale slopes, developing major landslide in the open pit Şırnak open pit coal mining. The rock fallings endangered the mining safety in recent years. This research provided stability patterns and cementing method strengthening cracks. The stages of experimentation explored the geo-technical characteristics and geoological formation. Fort his aim, four different open pit mining areas with similar geotechnical conditions, two main strengthening methods, and patterns of researches were developed. Firstly, data on landslides and rock dynamics over explosions were followed, and secondly, as happened commonly in the past, the same geological, geomorphologic, hydrological, climatic conditions were taken. Anchorage pile strengthenıng of slopes and cementing falling stone blocks were performed by mixtures of melted waste plastics/asphalt and fly ash for stability of higher slopes over 120 m height and over 65 degree in asphaltite mining site in Silopi and Avgamasya open pit No.1 mining site in Şırnak were carried out. On the other hand, due to that creep style rock falling from top of slopes, those melted polymer cementing of anchorage bolting and cracks, to eliminate those falling failure types and features, will be advantageous. The unconditional expectations related to this study was also defined for this region, such as the influence of the ground water, rock cracks and slope design, and explosion exchange dynamics leading to landslide. GEO5 software and manual stability analysis showed high risk area for plotting.

**Keywords:** mining pit, Şırnak asphaltite, active potential landslide, mining, geotechnical stability, slope stability

#### **1. Introduction**

Engineering geotechnical properties of surface units were determined by making geological map of Şırnak province and surrounding areas [1–5]. Geological mapping of slopes on large-scale topographic maps was one of the first landslide studies in the region. By determining the engineering properties of the slopes, it was aimed to draw attention to the importance of the area and the geotechnical criterion in the construction of the municipal development plans in the future.

Landslide is the downward slope due to the effects of massive soil masses or rocks on sloping slopes in mountainous areas such as gravity, slope, water, climatic factors, tectonics, weathering [6–9]. The geology of the material can be listed as precipitation, erosion, earthquake, and vegetation deprivation. Limit stress and balance analyses give accurate results in determining the landslide hazard and predicting future landslides [10].

In areas with high danger, landslide and related events will increase proportionally with increasing population density. It is very difficult to eliminate and reduce the risks arising from the processes of these landslides. There is an urgent need to better understand the character of the operation safety in open pit coal mining and to develop more predictive tools for stability.

Processes such as heavy rains, seismic, changes in groundwater level, erosion, climate, weathering, and natural topography are the natural parameters that trigger landslides. These effects increase the shear stress or decrease the shear resistance of the slope material [11]. Another important parameter that triggers landslide is urban activities. Increasing population and creating new living spaces forced people to settle on the slopes that present geological danger. The realization of equipment uses such as installing, practice, the creation of safe areas, and the realization of stress structure activities brought on by the excavation developed on the slopes can disrupt stability and create human activities and explosions that trigger landslides.

Especially in developing countries, the land in mountainous areas is not used in accordance with the topology, and wrong land use increases the probability of landslide development. Sustainability cannot be achieved in terms of physical environment, change and efficiency of landslide risk areas in open pit mining sites, and operation safety.

result of field and laboratory studies, and a topographic map has been created for all four hills by using the polar coordinate system. The high risk of rock sliding or falling damaging the asphaltite coal production occurred at the local open mining site 1 in Avgamasya seen as shown in **Figure 1b**. Black area was coal extracting area.

*(a and b) Satellite and contour topography of Avgamasya No.1 Pit Şırnak asphaltite coal mine site and survey*

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture…*

*DOI: http://dx.doi.org/10.5772/intechopen.91987*

Due to its tectonic structure and stratigraphy in the Southeastern Anatolia region, besides having reservoir rock and cover rock properties required for hydrological conditions, the water going down to the depths of the ground along the stretch cracks in the region are in a position to provide the necessary fluid for open pit asphaltite mining [31]. The crust of the earth was subjected to a stretching in the east-west direction with the effect of compression in the north-south direction throughout the region, and olivine basaltic magma rose from the asthenosphere

Basaltic magma, which reaches to the surface in the Karacadağ region between Diyarbakır-Şanlıurfa-Mardin, Gaziantep Yavuzeli region, and İdil-Cizre regions, flowed in several phases and left large areas under lava flows. Magma, which does not reach the earth, as in the north of Batman, created hot areas by making intrusions in several places. This situation in Şırnak is not clearly seen in the geological map of the region made by MTA General Directorate. The four lithostratigraphic units in the study area were distinguished from late period time age as Mardin

**2. The coal geology in Avgamasya, Şırnak**

along the stretch cracks formed.

**Figure 1.**

**111**

*area 1/18000 and 1/5000.*

Important landslides developed in open pit coal mining of the country in recent years are determined based on ground conditions. The stability of working area was managed with different methods such as geotechnical characters and geological formations, precautionary measurements, and the processes determined. Researchers have two basic theories for areas with similar geotechnical conditions [12–30]. Firstly, landslides are formed in the same geological, geomorphologic, hydrogeological, climatic conditions as in the past. Regarding the past phenomenon, the stability studies were carried out. Another is that the types and properties of landslides will be the same with other open pits. Therefore, knowing the mechanism and properties of past landslides is important basic information to evaluate landslides that may develop in the future, neighboring regions, or geotechnical similar areas. Geological and geotechnical analyzes of the slopes should be carried out in order to minimize the economic and social losses and casualties caused by landslides. In this direction, within the coal mining area of Şırnak Avgamasya and Silopi, open pit mining was carrying high landslide or rock falling risk (**Figure 1a**).

The geotechnical properties of the slopes where landslides occur in the districts that are 0.2–0.4 km from the south of the city and the center are analyzed, and the stability analyses were carried out with different methods using the GEO5 program. Within the scope of this project, a 1/5.000 scale engineering geology map covering 0.07 km2 of the study area and its surroundings that will be opened to urban use has been prepared as a *Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture… DOI: http://dx.doi.org/10.5772/intechopen.91987*

**Figure 1.**

**1. Introduction**

operation safety.

**110**

predicting future landslides [10].

*Mining Techniques - Past, Present and Future*

to develop more predictive tools for stability.

Engineering geotechnical properties of surface units were determined by making geological map of Şırnak province and surrounding areas [1–5]. Geological mapping of slopes on large-scale topographic maps was one of the first landslide studies in the region. By determining the engineering properties of the slopes, it was aimed to draw attention to the importance of the area and the geotechnical criterion

Landslide is the downward slope due to the effects of massive soil masses or rocks on sloping slopes in mountainous areas such as gravity, slope, water, climatic factors, tectonics, weathering [6–9]. The geology of the material can be listed as precipitation, erosion, earthquake, and vegetation deprivation. Limit stress and balance analyses give accurate results in determining the landslide hazard and

In areas with high danger, landslide and related events will increase proportionally with increasing population density. It is very difficult to eliminate and reduce the risks arising from the processes of these landslides. There is an urgent need to better understand the character of the operation safety in open pit coal mining and

Processes such as heavy rains, seismic, changes in groundwater level, erosion, climate, weathering, and natural topography are the natural parameters that trigger landslides. These effects increase the shear stress or decrease the shear resistance of the slope material [11]. Another important parameter that triggers landslide is urban activities. Increasing population and creating new living spaces forced people to settle on the slopes that present geological danger. The realization of equipment uses such as installing, practice, the creation of safe areas, and the realization of stress structure activities brought on by the excavation developed on the slopes can disrupt stability and create human activities and explosions that trigger landslides. Especially in developing countries, the land in mountainous areas is not used in

accordance with the topology, and wrong land use increases the probability of landslide development. Sustainability cannot be achieved in terms of physical environment, change and efficiency of landslide risk areas in open pit mining sites, and

Important landslides developed in open pit coal mining of the country in recent years are determined based on ground conditions. The stability of working area was managed with different methods such as geotechnical characters and geological formations, precautionary measurements, and the processes determined.

Researchers have two basic theories for areas with similar geotechnical conditions [12–30]. Firstly, landslides are formed in the same geological, geomorphologic, hydrogeological, climatic conditions as in the past. Regarding the past phenomenon, the stability studies were carried out. Another is that the types and properties of landslides will be the same with other open pits. Therefore, knowing the mechanism and properties of past landslides is important basic information to evaluate landslides that may develop in the future, neighboring regions, or geotechnical similar areas. Geological and geotechnical analyzes of the slopes should be carried out in order to minimize the economic and social losses and casualties caused by landslides. In this direction, within the coal mining area of Şırnak Avgamasya and Silopi,

open pit mining was carrying high landslide or rock falling risk (**Figure 1a**).

The geotechnical properties of the slopes where landslides occur in the districts that are 0.2–0.4 km from the south of the city and the center are analyzed, and the stability analyses were carried out with different methods using the GEO5 program. Within the scope of this project, a 1/5.000 scale engineering geology map covering 0.07 km2 of the study area and its surroundings that will be opened to urban use has been prepared as a

in the construction of the municipal development plans in the future.

*(a and b) Satellite and contour topography of Avgamasya No.1 Pit Şırnak asphaltite coal mine site and survey area 1/18000 and 1/5000.*

result of field and laboratory studies, and a topographic map has been created for all four hills by using the polar coordinate system. The high risk of rock sliding or falling damaging the asphaltite coal production occurred at the local open mining site 1 in Avgamasya seen as shown in **Figure 1b**. Black area was coal extracting area.

#### **2. The coal geology in Avgamasya, Şırnak**

Due to its tectonic structure and stratigraphy in the Southeastern Anatolia region, besides having reservoir rock and cover rock properties required for hydrological conditions, the water going down to the depths of the ground along the stretch cracks in the region are in a position to provide the necessary fluid for open pit asphaltite mining [31]. The crust of the earth was subjected to a stretching in the east-west direction with the effect of compression in the north-south direction throughout the region, and olivine basaltic magma rose from the asthenosphere along the stretch cracks formed.

Basaltic magma, which reaches to the surface in the Karacadağ region between Diyarbakır-Şanlıurfa-Mardin, Gaziantep Yavuzeli region, and İdil-Cizre regions, flowed in several phases and left large areas under lava flows. Magma, which does not reach the earth, as in the north of Batman, created hot areas by making intrusions in several places. This situation in Şırnak is not clearly seen in the geological map of the region made by MTA General Directorate. The four lithostratigraphic units in the study area were distinguished from late period time age as Mardin

Vulcanite (Upper Miocene) [30], Old Alluvium (Quaternary), New Alluvium (Quaternary), and Slope Rubble (Quaternary). Consisting of vulcanite, tuff, agglomerate and andesitic basaltic lava, syenitic rocks, which make up a large part of the study area, it shows as much shale and porous rock formations and slopes as occurred in the studied area chosen.

In recent years, both the opening of our university in our city and the migration movement from rural to provinces have also affected the coal economy in Şırnak. Open pit mining excavation and asphaltite production increased in different excavated pits in the area. A total of 500,000 asphaltite excavation by Asphaltite open pit mining with 20 separate pits continued to increase rapidly in Şırnak. Equipment and safety demands have also increased due to excessive extraction. This increase in demand caused people who are not competent in excavating in open pit mining production to enter the hauling, the control mechanism of dumping ability, modeling slopes to control the intensive stability, and areas that are not included in the development plan that are rapidly foreseen to mining development and safety. A vast majority of the excavatings in the Avgamasya open pit No. 1, 2, 3, 4 were excavated without adequate ground research. In the asphaltite mining of Şırnak, there were generally adjacent open pits in Silopi, Uludere. It is believed that such new open pits in Şırnak have been developed due to the neighboring hard condition of rock cracks, rock fallings, landslides, and sometimes the existing collapsed pits (or structural damage caused by the rotation of the equipment) or hydrological ground problems. Finally, Şırnak Avgamasya No. 2 landslide caused 8 workers' death, closing the excavation work. For these reasons, the investigation of the new buildings to be built in our province and the ground conditions on the basis of regional and parcels of new areas to be developed has become essential [18, 19]. The unconditional expectations related to this study were also defined for this region such as the influence of the ground water, rock cracks and slope design, and explosion exchange dynamics leading to landslide. GEO5 software and manual stability analysis showed high risk area for plotting (**Figure 2**).

degree in Uludere Ortabağ districts. Cizre and Merkez were areas of 2 degree risk. The studies considered to take precautions for steep slopes avoid those as below:

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture…*

• high slope type pits, excavating without basement control in real sense; and

• slope models designed without paying attention to hydrologic conditions and

Ground movements that may occur on a regional and/or 5–10 m basis have been observed in many irregular shale facing, high crack risk areas as seen in **Figure 4**.

*North and south steep slope faces of Avgamasya No. 1 pit of Şırnak asphaltite coal mine site and survey area.*

• steep slope structures over 60°;

*DOI: http://dx.doi.org/10.5772/intechopen.91987*

**Figure 3.**

**Figure 4.**

**113**

• earthquake resistant hollow-type structures;

*North face of Avgamasya No. 1 pit Şırnak asphaltite coal mine site and survey area.*

ground conditions did not escape attention.

Another issue for the province of Şırnak is the causes of the damages that may occur in possible explosions and earthquake shock waves caused by human conditions which are due to the engineering errors as well as the lack of control mechanism and construction defects in the excavating process. Şırnak Avgamasya open pit No. 1 mining site is located in the 2nd degree earthquake zone and these shortcomings mentioned above are also seen in our city. As seen in **Figure 3**, the risk of slope stability risk in explosions and excavation near earthquake district of 1

**Figure 2.** *Avgamasya No. 1 pit Şırnak asphaltite coal mine site and survey area.*

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture… DOI: http://dx.doi.org/10.5772/intechopen.91987*

**Figure 3.** *North face of Avgamasya No. 1 pit Şırnak asphaltite coal mine site and survey area.*

degree in Uludere Ortabağ districts. Cizre and Merkez were areas of 2 degree risk. The studies considered to take precautions for steep slopes avoid those as below:

• steep slope structures over 60°;

Vulcanite (Upper Miocene) [30], Old Alluvium (Quaternary), New Alluvium (Quaternary), and Slope Rubble (Quaternary). Consisting of vulcanite, tuff, agglomerate and andesitic basaltic lava, syenitic rocks, which make up a large part of the study area, it shows as much shale and porous rock formations and slopes as

stability analysis showed high risk area for plotting (**Figure 2**).

*Avgamasya No. 1 pit Şırnak asphaltite coal mine site and survey area.*

Another issue for the province of Şırnak is the causes of the damages that may occur in possible explosions and earthquake shock waves caused by human conditions which are due to the engineering errors as well as the lack of control mechanism and construction defects in the excavating process. Şırnak Avgamasya open pit No. 1 mining site is located in the 2nd degree earthquake zone and these shortcomings mentioned above are also seen in our city. As seen in **Figure 3**, the risk of slope stability risk in explosions and excavation near earthquake district of 1

In recent years, both the opening of our university in our city and the migration movement from rural to provinces have also affected the coal economy in Şırnak. Open pit mining excavation and asphaltite production increased in different excavated pits in the area. A total of 500,000 asphaltite excavation by Asphaltite open pit mining with 20 separate pits continued to increase rapidly in Şırnak. Equipment and safety demands have also increased due to excessive extraction. This increase in demand caused people who are not competent in excavating in open pit mining production to enter the hauling, the control mechanism of dumping ability, modeling slopes to control the intensive stability, and areas that are not included in the development plan that are rapidly foreseen to mining development and safety. A vast majority of the excavatings in the Avgamasya open pit No. 1, 2, 3, 4 were excavated without adequate ground research. In the asphaltite mining of Şırnak, there were generally adjacent open pits in Silopi, Uludere. It is believed that such new open pits in Şırnak have been developed due to the neighboring hard condition of rock cracks, rock fallings, landslides, and sometimes the existing collapsed pits (or structural damage caused by the rotation of the equipment) or hydrological ground problems. Finally, Şırnak Avgamasya No. 2 landslide caused 8 workers' death, closing the excavation work. For these reasons, the investigation of the new buildings to be built in our province and the ground conditions on the basis of regional and parcels of new areas to be developed has become essential [18, 19]. The unconditional expectations related to this study were also defined for this region such as the influence of the ground water, rock cracks and slope design, and explosion exchange dynamics leading to landslide. GEO5 software and manual

occurred in the studied area chosen.

*Mining Techniques - Past, Present and Future*

**Figure 2.**

**112**


Ground movements that may occur on a regional and/or 5–10 m basis have been observed in many irregular shale facing, high crack risk areas as seen in **Figure 4**.

**Figure 4.** *North and south steep slope faces of Avgamasya No. 1 pit of Şırnak asphaltite coal mine site and survey area.*

#### **3. Geotechnical properties of slopes and modeling**

#### **3.1 Field studies**

It is observed that the city center of Şırnak is located on a topography inclined to the south. Formations generally are composed of claystone and siltstone in the field.

It was observed and it is known that the center of Sirnak province is in Germav Formation. Germav Formation causes rapid formation of landslides as it forms a high slope topography by rapidly eroding due to its wear resistance. Therefore, in summary, Şırnak city center is located on the anchored Germav Formation, which is a mixture of sandy, carbonated, clayey, and silty units due to old landslides. Slope debris extends from the south of the stream (**Figure 2**) to the boundary of the working area. It has been determined with the field observations that the slope debris is composed of myocene limestones. Their thickness is extremely variable. They outcrop relatively less in areas where slope inclination decreases.

The surface of new alluvial deposits starting from city center to the south of Avgamasya open pit 1 (**Figure 3**) in the study area is gray marl shale. This part is generally covered with silty soil, and some parts of it consist of sandy and clayey zones. It has been determined that the new alluvium continues in the drilling up to 35 m by the Special Provincial Administration [32]. Slope Rubble is located south of the Stream in the study area.

Grain sizes range from fine clay to coarse sand. The thickness of the rubble, which does not show any grading and grading, varies between 10 and 35 cm. The active and potential landslide areas observed in the slope debris have been studied in detail.

#### **3.2 Geotechnical properties**

American Standards (ASTM 3080) were taken as a basis in the experiments carried out to determine the geotechnical properties of the surfaces surfaced in landslide risk areas. In the landslide risk areas of the study, undisturbed and lump rock samples were taken from the ground parts of the slope face. In experiments on the lines shown in **Figure 5**, the slope unit weight in wedge style and circle shape were considered in analysis with manual and GEO5 software and safety consistency limits were obtained.

The slope model construction plan is shown in **Figure 6**. The anchorage improved stability safety factors for steep sliding slopes in 20 m height excessive to 30 m. Wire mesh hanged top of slopes were avoiding rock falling of highly cracked shale block stones at 3–5 m size. The pile anchorage was designed and practiced for

hauling road slopes control and stability at the constructed near deep of slopes

*North and south steep slope faces of Avgamasya No. 1 pit of Şırnak asphaltite coal mine site and survey area*

*North and south steep slope faces of Avgamasya No. 1 pit of Şırnak asphaltite coal mine site and survey area*

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture…*

The uniaxial compressive test point load tests, RQD values, and Mohr-Coulomb criteria diagrammed were determined regarding ASTM standards; effective cohesion (c') and effective shear resistance friction angle (ϕ°)' of the rock samples were given in **Tables 1** and **2** on the standard rock samples of 50 mm cubic forms with the help

Undisturbed rock samples used in the point load experiment were classified according to ASTM standards and determined as cracked, altered ground. Rock samples show nonplastic character and there are also coarse-sized gravel.

In addition, during the making of these experiments, the unit volume weight,

The results obtained in plastic and liquid limit tests are given in **Table 1** for each sample. According to the ground classification in North district, the slope samples in the landslide N1 are in the less plastic shale and not plastic group, whereas in the region containing the landslide hazard N2 and N3, it is determined that there is less

the amount of compression, and the void ratio were determined.

bottom line as shown in **Figure 7**.

*with anchorage pile pattern.*

of ELE press equipment [23–25].

plastic.

**115**

**Figure 7.**

**Figure 6.**

*with anchorage pattern.*

*DOI: http://dx.doi.org/10.5772/intechopen.91987*

#### **Figure 5.**

*North and south steep slope faces of Avgamasya No. 1 pit of Şırnak asphaltite coal mine site and survey area with anchorage pattern.*

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture… DOI: http://dx.doi.org/10.5772/intechopen.91987*

**Figure 6.**

**3. Geotechnical properties of slopes and modeling**

*Mining Techniques - Past, Present and Future*

It is observed that the city center of Şırnak is located on a topography inclined to the south. Formations generally are composed of claystone and siltstone in the field. It was observed and it is known that the center of Sirnak province is in Germav Formation. Germav Formation causes rapid formation of landslides as it forms a high slope topography by rapidly eroding due to its wear resistance. Therefore, in summary, Şırnak city center is located on the anchored Germav Formation, which is a mixture of sandy, carbonated, clayey, and silty units due to old landslides. Slope debris extends from the south of the stream (**Figure 2**) to the boundary of the working area. It has been determined with the field observations that the slope debris is composed of myocene limestones. Their thickness is extremely variable.

They outcrop relatively less in areas where slope inclination decreases.

The surface of new alluvial deposits starting from city center to the south of Avgamasya open pit 1 (**Figure 3**) in the study area is gray marl shale. This part is generally covered with silty soil, and some parts of it consist of sandy and clayey zones. It has been determined that the new alluvium continues in the drilling up to 35 m by the Special Provincial Administration [32]. Slope Rubble is located south of

Grain sizes range from fine clay to coarse sand. The thickness of the rubble, which does not show any grading and grading, varies between 10 and 35 cm. The active and potential landslide areas observed in the slope debris have been studied

American Standards (ASTM 3080) were taken as a basis in the experiments carried out to determine the geotechnical properties of the surfaces surfaced in landslide risk areas. In the landslide risk areas of the study, undisturbed and lump rock samples were taken from the ground parts of the slope face. In experiments on the lines shown in **Figure 5**, the slope unit weight in wedge style and circle shape were considered in analysis with manual and GEO5 software and safety consistency

The slope model construction plan is shown in **Figure 6**. The anchorage improved stability safety factors for steep sliding slopes in 20 m height excessive to 30 m. Wire mesh hanged top of slopes were avoiding rock falling of highly cracked shale block stones at 3–5 m size. The pile anchorage was designed and practiced for

*North and south steep slope faces of Avgamasya No. 1 pit of Şırnak asphaltite coal mine site and survey area*

**3.1 Field studies**

the Stream in the study area.

**3.2 Geotechnical properties**

limits were obtained.

in detail.

**Figure 5.**

**114**

*with anchorage pattern.*

*North and south steep slope faces of Avgamasya No. 1 pit of Şırnak asphaltite coal mine site and survey area with anchorage pattern.*

#### **Figure 7.**

*North and south steep slope faces of Avgamasya No. 1 pit of Şırnak asphaltite coal mine site and survey area with anchorage pile pattern.*

hauling road slopes control and stability at the constructed near deep of slopes bottom line as shown in **Figure 7**.

The uniaxial compressive test point load tests, RQD values, and Mohr-Coulomb criteria diagrammed were determined regarding ASTM standards; effective cohesion (c') and effective shear resistance friction angle (ϕ°)' of the rock samples were given in **Tables 1** and **2** on the standard rock samples of 50 mm cubic forms with the help of ELE press equipment [23–25].

Undisturbed rock samples used in the point load experiment were classified according to ASTM standards and determined as cracked, altered ground. Rock samples show nonplastic character and there are also coarse-sized gravel.

In addition, during the making of these experiments, the unit volume weight, the amount of compression, and the void ratio were determined.

The results obtained in plastic and liquid limit tests are given in **Table 1** for each sample. According to the ground classification in North district, the slope samples in the landslide N1 are in the less plastic shale and not plastic group, whereas in the region containing the landslide hazard N2 and N3, it is determined that there is less plastic.


recompression parameters. In this case, parameters of the compacted soil can be

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture…*

In order to determine the slip resistance parameters of samples taken from different points of four separate slopes, a cutting box experiment was carried out. After the experiments, c<sup>0</sup> and ϕ<sup>0</sup> values were found. The stability analyses were carried out by the model constructed and practiced in the field as shown in

The manual weight chart method was so efficient and useful in slide pattern analysis in the area as shown in **Figure 9**. The sliding surface is circular half cylindrical. The sliding mass is divided into slices as equally as possible [33]. The safety coefficient values of the landslides were calculated according to

**Fellenius method**: if the forces between slices are considered to be in the same direction but opposite and equal to each other, they are not taken into account in the analysis. In the back, only the slice weight, ground reaction, cohesion, friction resistance, and leakage forces, if any, are in balance, without drainage on partially

The rupture envelope, which is determined by the strength's test under conditions, does not parallel the normal tension axis after a point, and the ground behaves both cohesively and internally. Total tension analysis method can be applied to cover this condition and to analyze stability by sliding mass divided into a certain number of vertical slices [12, 35–37]. **Bishop method**: in this method, as in all

*North and south steep slope faces of Avgamasya No.1 pit of Şırnak asphaltite coal mine site and survey area*

*with anchorage pile pattern and PE melting paste filling to cracks.*

used in stability analysis.

Fellenius, Bishop, and Jambu [1, 34].

*DOI: http://dx.doi.org/10.5772/intechopen.91987*

**Figures 8** and **9**.

water-saturated floors.

**Figure 8.**

**117**

**Table 1.**

*Results from geotechnical tests on samples taken from landslide slopes.*


**Table 2.**

*Proctor of ground samples and permeability test results.*

The results obtained from the geotechnical tests carried out on the samples taken from the slopes forming the landslide threat are given in **Tables 1** and **2**.

#### **4. Results and discussions**

The water content on the ground will be significantly affected by the clay content. When evaluated according to the percentage of clay in the floor, the floor samples show a non-cohesion or low cohesion feature.

The grain volume weights obtained by the experiments done on the samples taken from the landslide sites are shown in **Table 1**. In order to determine the soil types according to the grain size, grain distribution experiment has been carried out and the results and names in the combined soil classification are given in **Table 1**.

In order to determine the permeability of the ground, a fixed level permeability test instrument was used. The degree of permeability of the ground was determined by evaluating the results of the experiment (**Table 2**). When **Table 2** is examined, it is seen that the slopes S1, S2, and S3 fall under the permeable ground class.

The γ<sup>k</sup> and wopt values obtained as a result of Proctor experiments on soil samples taken from landslide areas are given in **Table 2**. With this experiment, optimum water content on the ground and maximum dry unit volume weight are determined and used for stability calculations of the slopes. Compaction parameters do not affect the stability of a natural slope, because these parameters are the parameters of the ground compacted in the desired way. In artificial slopes, compression parameters are used directly. If there is a landslide danger in a natural slope, in case of compression, the stabilization analyses are compared using these parameters. In the precautions to be taken against landslide danger, compacted filling can be made in front of the slope or bench slope can be made in the slope. At the same time, the natural ground is dug up and compacted according to the

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture… DOI: http://dx.doi.org/10.5772/intechopen.91987*

recompression parameters. In this case, parameters of the compacted soil can be used in stability analysis.

In order to determine the slip resistance parameters of samples taken from different points of four separate slopes, a cutting box experiment was carried out. After the experiments, c<sup>0</sup> and ϕ<sup>0</sup> values were found. The stability analyses were carried out by the model constructed and practiced in the field as shown in **Figures 8** and **9**.

The manual weight chart method was so efficient and useful in slide pattern analysis in the area as shown in **Figure 9**. The sliding surface is circular half cylindrical. The sliding mass is divided into slices as equally as possible [33].

The safety coefficient values of the landslides were calculated according to Fellenius, Bishop, and Jambu [1, 34].

**Fellenius method**: if the forces between slices are considered to be in the same direction but opposite and equal to each other, they are not taken into account in the analysis. In the back, only the slice weight, ground reaction, cohesion, friction resistance, and leakage forces, if any, are in balance, without drainage on partially water-saturated floors.

The rupture envelope, which is determined by the strength's test under conditions, does not parallel the normal tension axis after a point, and the ground behaves both cohesively and internally. Total tension analysis method can be applied to cover this condition and to analyze stability by sliding mass divided into a certain number of vertical slices [12, 35–37]. **Bishop method**: in this method, as in all

#### **Figure 8.**

*North and south steep slope faces of Avgamasya No.1 pit of Şırnak asphaltite coal mine site and survey area with anchorage pile pattern and PE melting paste filling to cracks.*

The results obtained from the geotechnical tests carried out on the samples taken

The water content on the ground will be significantly affected by the clay content. When evaluated according to the percentage of clay in the floor, the floor

The grain volume weights obtained by the experiments done on the samples taken from the landslide sites are shown in **Table 1**. In order to determine the soil types according to the grain size, grain distribution experiment has been carried out and the results and names in the combined soil classification are given in **Table 1**. In order to determine the permeability of the ground, a fixed level permeability test instrument was used. The degree of permeability of the ground was determined by evaluating the results of the experiment (**Table 2**). When **Table 2** is examined, it

is seen that the slopes S1, S2, and S3 fall under the permeable ground class. The γ<sup>k</sup> and wopt values obtained as a result of Proctor experiments on soil samples taken from landslide areas are given in **Table 2**. With this experiment, optimum water content on the ground and maximum dry unit volume weight are determined and used for stability calculations of the slopes. Compaction parameters do not affect the stability of a natural slope, because these parameters are the parameters of the ground compacted in the desired way. In artificial slopes, compression parameters are used directly. If there is a landslide danger in a natural slope, in case of compression, the stabilization analyses are compared using these parameters. In the precautions to be taken against landslide danger, compacted filling can be made in front of the slope or bench slope can be made in the slope. At the same time, the natural ground is dug up and compacted according to the

from the slopes forming the landslide threat are given in **Tables 1** and **2**.

samples show a non-cohesion or low cohesion feature.

**4. Results and discussions**

*Proctor of ground samples and permeability test results.*

**Rock formations**

**Table 1.**

**Table 2.**

**116**

γsat max (g/cm<sup>3</sup>

**Thickness (m)**

*Mining Techniques - Past, Present and Future*

**RQD (%)**

*Results from geotechnical tests on samples taken from landslide slopes.*

**c**0 **(kPa)** **φ**<sup>0</sup> **Pı (MPa)**

S1 25 25.9 3700 22 22.0 1.4 22 2.92 2.58 S2 74 42.9 3300 28 15.0 1.8 13 2.92 2.57 S3 25 40.8 2300 32 26.0 1.7 14 2.9 2.62 N1 47 25.9 2700 18 38.0 1.0 25 2.92 2.61 N2 55 35.4 4700 17 33.0 2.3 12 2.8 2.68 N3 46 33.9 4100 14 36.0 2.2 12 2.8 2.68

**Rock no S1 S2 S3 N1 N2 N3**

wopt (%) 15.9 11.9 11.0 12.3 3.8 3,3 Permeability (k) (cm/s) 5.3\*10–4 3.0\*10–5 6\*10–5 1.3\*10–4 3.3\*10–6 5.2\*10–6

) 2.98 2.75 2.77 2.98 2.75 2,77

**Iı (MPa; 50 mm)**

**Shear strength (mm/s)**

**γsat n (g/cm3 )**

**γdry (g/cm3 )**

sliding surfaces are more than those in the laboratories. It will be small, in other words, it will be closer to permanent values. In this respect, the value of 1.35 was

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture…*

According to the combined ground classification (USCS) located on the skirts of Şirnak Avgamasya open pit No.1 ground, movements in the area S1 consisting of shale group stones continued as high erosion creep with steep slope. Tectonic events occurring in the region have also triggered movements and continued to this day.

The discontinuous cracks were studied by cable extensometers weighted and observed as cm changes by daily periods in the control work. The slopes where landslide risk S1 is covered with pile anchorage and mesh slope stabilization by melted waste PE pressed (**Figures 8** and **10**). The maximum elevation difference between the top and heel point of the landslide S1 is 72 m, the maximum height of

In the observations made on this slope during the field studies, it was observed that it took the material that occurred in the stream and that small breaks and flows occurred after the precipitation. The geological map and landslide cross section and

Mineralogical compositions of the samples were determined by means of stan-

Prior to the preparation of the melting mixture by fly ash, micropictures showed the macropores in which melted PE penetrates into the pores and makes the pores stick easily to stabilize cracks as shown under the microscope pictures (**Figure 11**). Şirnak limestone, Midyat limestone, Şirnak marl, and marly shale were sized at 50–70 mm cubic blocks. The blocks were compressed with ELE press under 30 kN compression The blocks were crushed in test. The test results were given in

*Sliding shale landslide risk cross section with slope topography of Şırnak Avgamasya open pit No.1 mining,*

dard chemical Ca, Mg, and silica analyzes. The samples were first brought to dimensions between 40 mm and 10 mm in jaw crusher and were homogenized by milling to 0.1 mm. Powder samples are thawed and burned with HF in platinum crucible for silica content. Chemical composition of the rocks provided in the vicinity of Şirnak province in the experiments is given in **Table 3**. The amount of

**4.2 Slope S1 landslide risk analysis in Şırnak Avgamasya open pit No.1**

Nowadays, there are small-sized movements on the slope after the rains.

the surfaces where the calculations are made are given in **Figure 10**.

**4.3 Rock tests and crack stabilization by melted waste PE**

silica in the marly and marly limestone was reduced.

**Figure 10.**

**119**

*models for stability analysis.*

the slope is 80 m, and the slope angle of the slope is 58°.

taken as the limit security coefficient.

*DOI: http://dx.doi.org/10.5772/intechopen.91987*

**Figure 9.**

*Sliding shale areas contour topography of Şırnak Avgamasya open pit No.1 mining area and survey area 1/5000.*

stability problems, the initial shift is taken as post-equilibrium equations are taken out as if the slope is in the limit equilibrium. Bishop performed analysis with effective stresses instead of total stress. This method is more advanced than the methods brought by Taylor and Fellenius [28, 29]. **Janbu method**: this method can be used for all types of sliding surfaces, whether circular or not. In the slope stability analysis, it is a method that takes into account the inter-slice forces for the stability analysis of the more general types of noncircular shifts with the circular shifts occurring in the homogeneous split and fillings [38].

If the properties of the ground, very weak rock mass or rust material in any slope vary frequently throughout the slope, also the applicability of circular shear analysis methods disappears due to a structural feature such as ground-rock touch or in the presence of low shear strength planar levels in the mass [12, 13, 36, 37]. Shifts under such conditions: it develops along surfaces that start circularly in noncircular or near slope-top areas and continue in planar depths [14–17]. It is a method used to examine the stability of slopes, where instabilities develop or may develop along such sliding surfaces.

#### **4.1 Model slopes and landslide analysis**

In order to evaluate geological and landslide data, 1/1000 scale topographic maps of the slopes were prepared by field studies. Polar coordinate system was used in map production and Topcon GTS 702 electronic device was used. The heights are given according to the triangulation point on the 1200 m high hill. Topographic maps prepared for 4 slopes in the study area are given in **Figure 1**. In these 4 slopes, active and potential areas in terms of stability have been distinguished as a result of the surface studies on landslide.

According to the studies conducted, the regions where landslide is developed and the areas where relative movements are observed have been determined as active landslide area. Relative movements are determined by using the stress cracks on the surface. Potential landslide areas, on the other hand, correspond to areas where there are stress cracks around these active sites, but relative movements are not currently observed. In the study area, geological sections were prepared by marking the geological outcrops on topographic sections taken from 4 different slopes.

The safety coefficients of the slopes have been used with Bishop, Janbu, and Fellenius methods and circular slip diagrams GEO5 program for different slip surfaces [39, 40]. It is taken as 1.3 in ASTM standard in the limitations of safety coefficients. In laboratory experiments, c<sup>0</sup> and φ<sup>0</sup> were found according to the effective and maximum resistance parameters. It is known that movements occur in the stretch cracks in the field over time. In this case, the internal parameters on the

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture… DOI: http://dx.doi.org/10.5772/intechopen.91987*

sliding surfaces are more than those in the laboratories. It will be small, in other words, it will be closer to permanent values. In this respect, the value of 1.35 was taken as the limit security coefficient.

#### **4.2 Slope S1 landslide risk analysis in Şırnak Avgamasya open pit No.1**

According to the combined ground classification (USCS) located on the skirts of Şirnak Avgamasya open pit No.1 ground, movements in the area S1 consisting of shale group stones continued as high erosion creep with steep slope. Tectonic events occurring in the region have also triggered movements and continued to this day. Nowadays, there are small-sized movements on the slope after the rains.

The discontinuous cracks were studied by cable extensometers weighted and observed as cm changes by daily periods in the control work. The slopes where landslide risk S1 is covered with pile anchorage and mesh slope stabilization by melted waste PE pressed (**Figures 8** and **10**). The maximum elevation difference between the top and heel point of the landslide S1 is 72 m, the maximum height of the slope is 80 m, and the slope angle of the slope is 58°.

In the observations made on this slope during the field studies, it was observed that it took the material that occurred in the stream and that small breaks and flows occurred after the precipitation. The geological map and landslide cross section and the surfaces where the calculations are made are given in **Figure 10**.

#### **4.3 Rock tests and crack stabilization by melted waste PE**

Mineralogical compositions of the samples were determined by means of standard chemical Ca, Mg, and silica analyzes. The samples were first brought to dimensions between 40 mm and 10 mm in jaw crusher and were homogenized by milling to 0.1 mm. Powder samples are thawed and burned with HF in platinum crucible for silica content. Chemical composition of the rocks provided in the vicinity of Şirnak province in the experiments is given in **Table 3**. The amount of silica in the marly and marly limestone was reduced.

Prior to the preparation of the melting mixture by fly ash, micropictures showed the macropores in which melted PE penetrates into the pores and makes the pores stick easily to stabilize cracks as shown under the microscope pictures (**Figure 11**).

Şirnak limestone, Midyat limestone, Şirnak marl, and marly shale were sized at 50–70 mm cubic blocks. The blocks were compressed with ELE press under 30 kN compression The blocks were crushed in test. The test results were given in

#### **Figure 10.**

*Sliding shale landslide risk cross section with slope topography of Şırnak Avgamasya open pit No.1 mining, models for stability analysis.*

stability problems, the initial shift is taken as post-equilibrium equations are taken out as if the slope is in the limit equilibrium. Bishop performed analysis with effective stresses instead of total stress. This method is more advanced than the methods brought by Taylor and Fellenius [28, 29]. **Janbu method**: this method can be used for all types of sliding surfaces, whether circular or not. In the slope stability analysis, it is a method that takes into account the inter-slice forces for the stability analysis of the more general types of noncircular shifts with the circular shifts

*Sliding shale areas contour topography of Şırnak Avgamasya open pit No.1 mining area and survey area*

If the properties of the ground, very weak rock mass or rust material in any slope vary frequently throughout the slope, also the applicability of circular shear analysis methods disappears due to a structural feature such as ground-rock touch or in the presence of low shear strength planar levels in the mass [12, 13, 36, 37]. Shifts under such conditions: it develops along surfaces that start circularly in noncircular or near slope-top areas and continue in planar depths [14–17]. It is a method used to examine the stability of slopes, where instabilities develop or may develop along

In order to evaluate geological and landslide data, 1/1000 scale topographic maps of the slopes were prepared by field studies. Polar coordinate system was used in map production and Topcon GTS 702 electronic device was used. The heights are given according to the triangulation point on the 1200 m high hill. Topographic maps prepared for 4 slopes in the study area are given in **Figure 1**. In these 4 slopes, active and potential areas in terms of stability have been distinguished as a result of

According to the studies conducted, the regions where landslide is developed and the areas where relative movements are observed have been determined as active landslide area. Relative movements are determined by using the stress cracks on the surface. Potential landslide areas, on the other hand, correspond to areas where there are stress cracks around these active sites, but relative movements are not currently observed. In the study area, geological sections were prepared by marking the geological outcrops on topographic sections taken from 4 different

The safety coefficients of the slopes have been used with Bishop, Janbu, and Fellenius methods and circular slip diagrams GEO5 program for different slip surfaces [39, 40]. It is taken as 1.3 in ASTM standard in the limitations of safety coefficients. In laboratory experiments, c<sup>0</sup> and φ<sup>0</sup> were found according to the effective and maximum resistance parameters. It is known that movements occur in the stretch cracks in the field over time. In this case, the internal parameters on the

occurring in the homogeneous split and fillings [38].

**4.1 Model slopes and landslide analysis**

*Mining Techniques - Past, Present and Future*

the surface studies on landslide.

slopes.

**118**

such sliding surfaces.

**Figure 9.**

*1/5000.*


#### **Table 3.**

*The chemical analysis of rock specimens of Şırnak asphaltite Avgamasya open pit mine No.1 site limestone, marly shale, and shale.*

**Figure 11.** *(a) Shale, (b) marly shale, (c) Şırnak porous limestone, and (d) the Şırnak-altered porous limestone.*

#### **Figures 12** and **13** within the limit indentation pattern values charted, as given in **Table 3** (**Figures 14**–**16**).

The rock samples have also increased the 13.6% of the pore contained in bulk pile as 25 mm fraction showed a pore rate as lower at 9.2%. The results were shown in **Figure 1**. **Figure 1** also proved that the melted interaction of 20 min is sufficient. Hence, the dissolution process reached the PE and ash saturation of the emulsified PE solution by waste machine oil. The pores in the limestone sample reached 13.6%

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture…*

(**Table 4**).

**121**

**Figure 13.**

**Figure 12.**

*Compression strength change of the Şırnak aggregates.*

*DOI: http://dx.doi.org/10.5772/intechopen.91987*

*Point load strength change of the Şırnak aggregates.*

#### **4.4 Point loading and compressive strength tests of rocks**

The test samples were produced as 5 5 5 cm blocks and 10 samples were determined to be with 95% accuracy rate by prestigious ELE brand test equipment. The results are shown in **Figures 6** and **7**.

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture… DOI: http://dx.doi.org/10.5772/intechopen.91987*

**Figure 12.** *Compression strength change of the Şırnak aggregates.*

**Figure 13.** *Point load strength change of the Şırnak aggregates.*

The rock samples have also increased the 13.6% of the pore contained in bulk pile as 25 mm fraction showed a pore rate as lower at 9.2%. The results were shown in **Figure 1**. **Figure 1** also proved that the melted interaction of 20 min is sufficient. Hence, the dissolution process reached the PE and ash saturation of the emulsified PE solution by waste machine oil. The pores in the limestone sample reached 13.6% (**Table 4**).

**Figures 12** and **13** within the limit indentation pattern values charted, as given in

*(a) Shale, (b) marly shale, (c) Şırnak porous limestone, and (d) the Şırnak-altered porous limestone.*

**%Component Sırnak marly shale stone Şırnak Marl Şırnak porous limestone Şırnak shale** SiO2 9.42 24.14 2.12 48.53 Al2O3 6.53 12.61 1.71 24.61 Fe2O3 4.48 7.34 0.58 7.59 CaO 39.23 29.18 45.22 9.48 MgO 2.28 4.68 7.41 3.28 K2O 0.53 3.32 0.40 2.51 Na2O 0.24 1.11 0.21 0.35 Ignition loss 26.11 21.43 48.04 3.09 SO3 0.21 0.20 0.02 0.32

*The chemical analysis of rock specimens of Şırnak asphaltite Avgamasya open pit mine No.1 site limestone,*

The test samples were produced as 5 5 5 cm blocks and 10 samples were determined to be with 95% accuracy rate by prestigious ELE brand test equipment.

**4.4 Point loading and compressive strength tests of rocks**

**Table 3** (**Figures 14**–**16**).

**Figure 11.**

**120**

**Table 3.**

*marly shale, and shale.*

*Mining Techniques - Past, Present and Future*

The results are shown in **Figures 6** and **7**.

#### **Figure 14.**

*Histogram view of impact, abrasion resistance: 1. Sırnak porous limestone; 2. Sırnak marl; 3. Şırnak marly shale; 4. Midyat limestone; 5. Şırnak porous limestone; and 6. Şırnak marl.*

microporous structure of 13.4–14.8% silica in Sırnak marly limestone and as 5–30 micron microporous. The degree of chemical interaction in marinated limestone and marl was not sufficient due to the silica content and the microcrystalline pore

*Chemical composition of binder fillers for crack stabilization in the Şırnak asphaltite Avgamasya open pit*

**%Component Shale powder Asphaltite slime Tatvan pumice Şırnak fly ash** SiO2 43.48 50.50 60.13 41.48 Al2O3 16.10 14.61 17.22 18.10 Fe2O3 10.52 24.30 4.59 4.52 CaO 8.48 2.30 2.48 18.48 MgO 3.80 1.28 2.17 4.20 K2O 2.51 2.51 3.51 2.71 Na2O 1.35 1.35 4.35 1.95 Ign.loss. 10.9 0.21 4.12 1.9 SO3 0.32 0.12 0.52 0.22

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture…*

*Open pit mine No.2: excavations for steep slopes—asphaltite excavation.*

*DOI: http://dx.doi.org/10.5772/intechopen.91987*

According to the information obtained from the drillings carried out by the Special Provincial Administration, the depth of the slope varies 11–20 m. In the investigations, the groundwater level on the slope where the landslide no S1 is located was observed at a base of 25 m. In the stability calculations made on the

separately according to "saturated" rock were shown in **Table 6**. According to Hoek-Bray [32], the safety coefficient value is determined as 1.25. When **Table 7** is

. Safety coefficient values of possible sliding surfaces calculated

, φ<sup>0</sup> = 22°, γnatural = 2.97 g/cm<sup>3</sup>

,

structure.

**Table 4.**

**Figure 16.**

*mine No.1.*

and γdry = 2.7 g/cm<sup>3</sup>

**123**

**4.5 Stability risk survey for S1 slope**

slope where landslide S1 developed, c<sup>0</sup> = 190 kg/cm<sup>2</sup>

**Figure 15.** *Chart pattern for Workers in open pit mine excavations for S1 slope. Excavation was completed till 43° slope obtained for overburden.*

Porous limestone texture, chemical interaction result, and petrographic changes are seen in **Figure 2**. It was determined that the amount of silicate of 2.3 and 4% of Şirnak altered limestone was changed and the pore size was as micro mesh with a size of 1–3 mm macro 50–300 micron. This pore amount is in microcrystalline

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture… DOI: http://dx.doi.org/10.5772/intechopen.91987*

**Figure 16.**

*Open pit mine No.2: excavations for steep slopes—asphaltite excavation.*


#### **Table 4.**

*Chemical composition of binder fillers for crack stabilization in the Şırnak asphaltite Avgamasya open pit mine No.1.*

microporous structure of 13.4–14.8% silica in Sırnak marly limestone and as 5–30 micron microporous. The degree of chemical interaction in marinated limestone and marl was not sufficient due to the silica content and the microcrystalline pore structure.

#### **4.5 Stability risk survey for S1 slope**

According to the information obtained from the drillings carried out by the Special Provincial Administration, the depth of the slope varies 11–20 m. In the investigations, the groundwater level on the slope where the landslide no S1 is located was observed at a base of 25 m. In the stability calculations made on the slope where landslide S1 developed, c<sup>0</sup> = 190 kg/cm<sup>2</sup> , φ<sup>0</sup> = 22°, γnatural = 2.97 g/cm<sup>3</sup> , and γdry = 2.7 g/cm<sup>3</sup> . Safety coefficient values of possible sliding surfaces calculated separately according to "saturated" rock were shown in **Table 6**. According to Hoek-Bray [32], the safety coefficient value is determined as 1.25. When **Table 7** is

Porous limestone texture, chemical interaction result, and petrographic changes are seen in **Figure 2**. It was determined that the amount of silicate of 2.3 and 4% of Şirnak altered limestone was changed and the pore size was as micro mesh with a size of 1–3 mm macro 50–300 micron. This pore amount is in microcrystalline

*Chart pattern for Workers in open pit mine excavations for S1 slope. Excavation was completed till 43° slope*

*Histogram view of impact, abrasion resistance: 1. Sırnak porous limestone; 2. Sırnak marl; 3. Şırnak marly*

*shale; 4. Midyat limestone; 5. Şırnak porous limestone; and 6. Şırnak marl.*

*Mining Techniques - Past, Present and Future*

**Figure 14.**

**Figure 15.**

**122**

*obtained for overburden.*

examined, since the boundary safety coefficient is taken as 1.25, it can be easily seen that the sliding slopes S1 and S2 were unstable. It is seen that the sliding surfaces 3 and 4 are very close to the limit value when it is made according to "saturated" rock and it is unstable when it is examined according to "natural."

the fracture distribution percentage and angle of the slope angle in the formation,

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture…*

The fracture or discontinuity angle t frequency% in the 20 m sliding on slope

R*i*Fi tan θ ¼

**Compacted binder 1. Sample 2. Sample 3. Sample 4. Sample Average sample**

 205,6 205,6 209,60 203,5 204,55 284,5 284,5 288,4 296,5 289,80 15% fly ash 182,99 205,6 188,81 198,39 196,1 194,43 284,5 304,84 319,7 301,75 308,76 20% fly ash 203,61 205,6 251,17 253,9 257,56 254,21 284,5 360,79 369,2 360,73 363,57 25% fly ash 205,50 205,6 300,24 220,92 241,46 254,21 284,5 355,13 356,99 359,21 357,11 30% fly ash 281,89 205,6 323,66 316,99 313,72 318,12 284,5 381,17 377,42 371,8 376,80

**Site F1 F2 F3 F4 F5 F6 F7 F8 F9 F10 SF** S1 1210 2150 3270 4180 5460 6210 5740 4130 3550 2230 1.1 S2 1320 2280 3330 4270 5640 6540 5730 4580 3840 2470 1.2 S3 1110 2010 3110 4080 5200 6030 5530 3970 3110 1940 1.3 N1 1200 2100 3200 4100 5400 6200 5700 4100 3500 2200 1.2 N2 1200 2100 3200 4100 5400 6200 5700 4100 3500 2200 1.25 N3 1200 2100 3200 4100 5400 6200 5700 4100 3500 2200 1.3

*Sliding shale landslide risk cross section with slope topography of Şırnak Avgamasya open pit No.1 mining.*

�*tiθ*

ð*b a e* �*tiθ*

*dx* were calculated as given in

*dy<sup>i</sup>* (3)

**)**

*dy* (4)

which has a high probability of slide.

*DOI: http://dx.doi.org/10.5772/intechopen.91987*

below equations and in **Tables 5**–**7**

5% fly ash

**Table 5.**

**Table 6.**

**125**

*Şırnak limestone aggregate sieve analysis results aggregate type.*

*5 m unit saturated weight charts kN.*

direction and the variable position in the design card *dy*

<sup>R</sup>*<sup>c</sup>* <sup>¼</sup> <sup>X</sup> i

0

**Curing time (min) Uniaxial compressive strength (kg/cm2**

*dy dx* <sup>¼</sup> *<sup>e</sup>*

According to the slope-clay floor combined ground classification (USCS) located in the upper skirts of the south district, the SW-SC group consists of plastic floors. Tectonic events occurring in the region have also triggered movements and continued to this day. Today, it has been observed in the field studies that there are small size movements on the slope after the rains. The slope where landslide risk S1 was developed was covered with color black showed the instability (**Figure 11**). The maximum elevation difference between the top and heel point of the landslide S1 is 75 m, the maximum height of the slope is 110 m, and the slope angle of the slope is 48°.

According to the information obtained from the drillings carried out by the Mining Bureau of Şirnak Administration Authority, the thickness of the slope varies between 11 and 20 m. In the investigations, underground water level was observed on the slope where the landslide number. In the stability calculations on the slope where the landslide S1 is developed, c<sup>0</sup> = 199 kg/ cm<sup>2</sup> , φ <sup>0</sup> = 22°, γsat = 2.97 g/ cm<sup>3</sup> , and γdry = 2.77 g/cm<sup>3</sup> were used. Safety coefficient values of possible sliding surfaces calculated separately according to "saturated" rocks were shown in **Table 6**. According to Hoek-Bray, the safety coefficient value is determined as 1.25. When **Tables 6** and **7** were examined and analyzed, since the boundary safety coefficient was taken as 1.25, it could be easily seen that the sliding surfaces 1 and 2 are unstable. It was seen that the sliding surfaces S3 and S4 are very close to the limit value according to saturated rock weight.

Shale rocks without drainage develop geotechnical parameters anisotropically. The slopes under compressive load show different shear strength depending on the shear face direction depending on the water content.

$$c\_{\theta} = c\_{2} + (c\_{1} - c\_{2}) \cos^{2}\theta \tag{1}$$

where *θ* is the angle made with the principal stress direction where shear stress occurs as below:


Compressive load values in the calculation of design model slope stability cards were given as equation below:

$$F = \sum\_{0}^{i} \mathbf{N}\_{i} \mathbf{F}\_{i} = \mathbf{N}\_{i} \frac{C\_{i}}{\gamma H} \cos \rho^{i} \tag{2}$$

The compression load on the N*i*block was calculated as the ground anisotropic shale shear *<sup>C</sup>*<sup>1</sup> *Ci* strength values depending on the slope angle, *β*. Although safety factor is designed as 1.25 and 1.35, safety factor is preferred as 1.35, according to water content greatly considered.

Due to the differences in fracture distribution, in order to determine the safe slope stability angle in the stress design cards, the sliding risk factor R*<sup>c</sup>* depends on *Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture… DOI: http://dx.doi.org/10.5772/intechopen.91987*

the fracture distribution percentage and angle of the slope angle in the formation, which has a high probability of slide.

The fracture or discontinuity angle t frequency% in the 20 m sliding on slope direction and the variable position in the design card *dy dx* were calculated as given in below equations and in **Tables 5**–**7**

$$\mathbf{R}\_{\mathbf{c}} = \sum\_{0}^{\mathrm{i}} \mathbf{R}\_{\mathbf{i}} \mathbf{F}\_{\mathbf{i}} \tan \theta = \int\_{a}^{b} e^{-ti\theta} d\mathbf{y}^{\mathrm{i}} \tag{3}$$

$$\frac{dy}{dx} = e^{-\text{ti}\theta} dy\tag{4}$$


**Table 5.**

examined, since the boundary safety coefficient is taken as 1.25, it can be easily seen that the sliding slopes S1 and S2 were unstable. It is seen that the sliding surfaces 3 and 4 are very close to the limit value when it is made according to "saturated" rock

According to the slope-clay floor combined ground classification (USCS) located in the upper skirts of the south district, the SW-SC group consists of plastic floors. Tectonic events occurring in the region have also triggered movements and

continued to this day. Today, it has been observed in the field studies that there are small size movements on the slope after the rains. The slope where landslide risk S1 was developed was covered with color black showed the instability (**Figure 11**). The maximum elevation difference between the top and heel point of the landslide S1 is 75 m, the maximum height of the slope is 110 m, and the slope angle of the

According to the information obtained from the drillings carried out by the Mining Bureau of Şirnak Administration Authority, the thickness of the slope varies between 11 and 20 m. In the investigations, underground water level was observed on the slope where the landslide number. In the stability calculations on the slope

γdry = 2.77 g/cm<sup>3</sup> were used. Safety coefficient values of possible sliding surfaces calculated separately according to "saturated" rocks were shown in **Table 6**. According to Hoek-Bray, the safety coefficient value is determined as 1.25. When **Tables 6** and **7** were examined and analyzed, since the boundary safety coefficient was taken as 1.25, it could be easily seen that the sliding surfaces 1 and 2 are unstable. It was seen that the sliding surfaces S3 and S4 are very close to the limit

Shale rocks without drainage develop geotechnical parameters anisotropically. The slopes under compressive load show different shear strength depending on the

*<sup>c</sup><sup>θ</sup>* <sup>¼</sup> *<sup>c</sup>*<sup>2</sup> <sup>þ</sup> ð Þ *<sup>c</sup>*<sup>1</sup> � *<sup>c</sup>*<sup>2</sup> cos <sup>2</sup>

where *θ* is the angle made with the principal stress direction where shear stress

Compressive load values in the calculation of design model slope stability cards

*C*i

N*i*Fi ¼ N*<sup>i</sup>*

The compression load on the N*i*block was calculated as the ground anisotropic

factor is designed as 1.25 and 1.35, safety factor is preferred as 1.35, according to

Due to the differences in fracture distribution, in order to determine the safe slope stability angle in the stress design cards, the sliding risk factor R*<sup>c</sup>* depends on

*Ci* strength values depending on the slope angle, *β*. Although safety

, φ <sup>0</sup> = 22°, γsat = 2.97 g/ cm<sup>3</sup>

*θ* (1)

*<sup>γ</sup><sup>H</sup>* cos *<sup>β</sup><sup>i</sup>* (2)

, and

and it is unstable when it is examined according to "natural."

*Mining Techniques - Past, Present and Future*

where the landslide S1 is developed, c<sup>0</sup> = 199 kg/ cm<sup>2</sup>

shear face direction depending on the water content.

• compressive strength in the direction of θ sliding face;

*<sup>F</sup>* <sup>¼</sup> <sup>X</sup> i

0

• normal load stress in the vertical direction; and

• shear stress in the horizontal direction.

were given as equation below:

water content greatly considered.

value according to saturated rock weight.

slope is 48°.

occurs as below:

shale shear *<sup>C</sup>*<sup>1</sup>

**124**

*Şırnak limestone aggregate sieve analysis results aggregate type.*


#### **Table 6.**

*Sliding shale landslide risk cross section with slope topography of Şırnak Avgamasya open pit No.1 mining.*


**Table 7.** *Sliding shale landslide risk finite analysis with slope topography of Şırnak Avgamasya open pit No.1 mining.*

Safety risk parameter was calculated as 1.42 stable for 400 slopes, but for 500 and 600 slopes, the safety factors decreased to 1198 and 1060. As given in the figure, the equation slope 44.20 has been given a safety factor for a stable slope as

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture…*

In the observations made on this slope during the field studies, it was observed that it took the material that occurred in the stream and that small breaks and flows occurred after the precipitation. The geological map and

and development of nonslip methods in the region is of special importance

landslide cross section and the surfaces where the calculations are made are given in

Groundwater movement change should be considered as a hazard and landslide prevention methods appropriate for the site should be determined. In addition, as the study area will be opened to urban use within the scope of the project, research

ð5Þ

1320 is shown in **Figure 8**.

*DOI: http://dx.doi.org/10.5772/intechopen.91987*

**Figure 17**.

(**Tables 8**–**10**).

**Chart Block height**

**Figure 17.**

**Table 8.**

**127**

*Weight chart calculations for S11.*

**Block width**

*Sliding shale landslide risk isocontour map; black area is unsafe area.*

**Block weight (ton)**

1 3 4 11,12 109,0872 6,217,489 2 5 6 27,8 272,718 13,29,372 3 10 8 74,13,333 727,248 32,94,993 4 15 9 125,1 1227,231 54,57,176 5 18 7 116,76 1145,416 51,03364 6 16 5 74,13,333 727,248 32,94,993 7 14 4 51,89,333 509,0736 23,51,495 8 11 4 40,77,333 399,9864 18,79,746 9 9 3 25,02 245,4462 12,11,435

10 7 3 19,46 190,9026 9,755,607 500 Total 5554,357 255,1988 1206

**Block weight (kN)**

**Block chart share (MPa)**

**Safety**

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture… DOI: http://dx.doi.org/10.5772/intechopen.91987*

Safety risk parameter was calculated as 1.42 stable for 400 slopes, but for 500 and 600 slopes, the safety factors decreased to 1198 and 1060. As given in the figure, the equation slope 44.20 has been given a safety factor for a stable slope as 1320 is shown in **Figure 8**.

$$S = \frac{\Sigma(c' + \sigma' \cdot \tan \phi') \cdot \ell}{\Sigma(W \cdot \sin \phi')} \tag{5}$$

In the observations made on this slope during the field studies, it was observed that it took the material that occurred in the stream and that small breaks and flows occurred after the precipitation. The geological map and landslide cross section and the surfaces where the calculations are made are given in **Figure 17**.

Groundwater movement change should be considered as a hazard and landslide prevention methods appropriate for the site should be determined. In addition, as the study area will be opened to urban use within the scope of the project, research and development of nonslip methods in the region is of special importance (**Tables 8**–**10**).

**Figure 17.** *Sliding shale landslide risk isocontour map; black area is unsafe area.*


#### **Table 8.** *Weight chart calculations for S11.*

**Site**

**126**

S1

S2

S3

N1

N2

N3

*5 m unit saturated rock block weight points kN (H/V).*

**Table 7.** *Sliding shale landslide risk finite analysis with slope topography*

 *of Şırnak Avgamasya*

 *open pit No.1 mining.*

 321/332

 515/443

 527/453

 418/366

 546/377

 621/556

 574/221

 513/462

 655/555

 223/112

 418/366

 513/462

 655/555

 1.3

 321/332

 515/443

 527/453

 418/366

 546/377

 621/556

 574/221

 513/462

 655/555

 223/112

 418/366

 513/462

 655/555

 1.25

 321/332

 515/443

 527/453

 418/366

 546/377

 621/556

 574/221

 513/462

 655/555

 223/112

 418/366

 513/462

 655/555

 1.2

 321/332

 515/443

 527/453

 418/366

 546/377

 621/556

 574/221

 513/462

 655/555

 223/112

 418/366

 513/462

 655/555

 1.3

 321/332

 515/443

 527/453

 418/366

 546/377

 621/556

 574/221

 513/462

 655/555

 223/112

 418/366

 513/462

 655/555

 1.2

 321/332

 515/443

 527/453

 418/366

 546/377

 621/556

 574/221

 513/462

 655/555

 223/112

 418/366

 513/462

 655/555

 1.1

 **Joint 1,**

 **Joint 2**

 **Joint 3**

 **Joint 4**

 **Joint 5**

 **Joint 6**

 **Joint 7**

 **Joint 8**

 **Joint 9 Joint 10**

 **Joint 11**

 **Joint 12**

 **Joint 13**

 **SF 13-12**

*Mining Techniques - Past, Present and Future*


In order to prevent this situation with clay, it is necessary to accumulate rock material in sizes that the river cannot carry, or the creek must be improved. South and North districts of the Avgamasya open pit No.1 asphaltite mine should be reinforced in the hazardous areas according to ground water. As a result of these shock wave movements, taking precautionary measures according to the slopes and underground water floods, drainage channels should be examined again. For this reason, there is a need to investigate the stability of the slopes that will be opened to

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture…*

The importance of extensometer station was very critical for over 20 mm displacement and covering the cracks by mixture of melted plastics and asphalt and facilitates the infiltration of rainwater into the channel reduces the landslide or rock

which are positive in terms of stability, and hence, there is a reduction in the holding forces that keep the slopes in balance. For this reason, enrichment in terms of vegetation is an important landslide preventing parameter in the region. However, the effect of vegetation on stability will be minimal for sliding surfaces up to 30 m in depth. The weathering causes the rocks to change to a great extent, the bond between the grains to weaken and disappear completely. The rocks, which weaken as a result of weathering in the study area, are easily eroded and change the slope angles and slope altitudes. The weathering observed in the rocks in the study area also contributes negatively to the stability problems. As a result of geotechnical analysis carried out in the study area, it is concluded that very large landslides will not be expected in the future. However, this result does not eliminate the possibility of landslide danger for urban areas and urban development areas. For this reason, especially in large-scale plans such as master zoning plan and application zoning plan, geological hazard in the region should be evaluated and landslide prevention methods appropriate for the site should be determined. In addition, research and development methods to prevent instability in the region are of special importance since the study area will be opened to urban use within the scope of the project. As a result of laboratory experiments carried out on the soil samples, it was determined that the slope material was permeable, the cohesion value ranged between 1300 and 3800 kPa, and the internal friction angle ranged between 37.5 and 22.1, and according to the combined ground classification, the slope material consisted of Shale, low marly shale, and cracked limestone low RQD group rocks. From the stability analysis performed in the light of this information, it was concluded that the slopes S1, S2, and N1 were unstable, and the slopes S3 and N2 N3

The lack of vegetation in the study area prevents benefiting from these effects,

As a result of the field investigations, it was determined that the increase in the

Plants facilitate the infiltration of rainwater into the mass and slow down and reduce superficial flow. This prevents the masses from erosion. The roots of the

flow rate at the peak with the melting of snow in May and June caused severe erosion on the heel of the slopes and thus had a negative effect on the stability of the slope. In order to prevent this situation, which is effective on slopes N2 and N3, it is necessary to accumulate rock material in sizes that the stream cannot bear, or the stream should be improved. Sirnak City and the surrounding area, according to Turkey Earthquake Zone Map, are located in the first degree in the danger zone. Since these regions are within the influenced area of the South East Anatolian Fault, frequent earthquakes occur in the region and some tectonic movements occur due to these earthquakes. As a result of these movements, the stability of the slopes is compromised. For this reason, there is a need to investigate the stability of the

flow. This drainage patterns prevents the masses creep.

*DOI: http://dx.doi.org/10.5772/intechopen.91987*

urban use.

were stable.

**129**

slopes that will be opened to urban use.

plants, whose roots reach deep, are mechanical.

#### **Table 9.**

*Weight chart calculations for S12.*


**Table 10.** *Weight chart calculations for S13.*

#### **5. Conclusions**

As a result of the laboratory experiments carried out on the soil samples, it was determined that the slope material was permeable, the cohesion value ranged between 0.1 and 0.38, the internal friction angle ranged between 17.5 and 22.4, and according to the combined ground classification, the slope material consisted of shale group plastic steep slopes in the top area of low RQD. From the stability analysis performed in the light of this information, it was concluded that the slopes in S1, S2, and N1 are unstable, and the slope S3 and N2 in open pit were stable.

As a result of the field studies, it was determined that the increase in the flow rate at the peak with the melting of snow, especially in May and June, caused severe erosion on the heel of the slopes and thus had a negative effect on the stability of the slope. Chart patterns were obtained on the slopes S1, S2, and S3 to reduce sliding risk. *Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture… DOI: http://dx.doi.org/10.5772/intechopen.91987*

In order to prevent this situation with clay, it is necessary to accumulate rock material in sizes that the river cannot carry, or the creek must be improved. South and North districts of the Avgamasya open pit No.1 asphaltite mine should be reinforced in the hazardous areas according to ground water. As a result of these shock wave movements, taking precautionary measures according to the slopes and underground water floods, drainage channels should be examined again. For this reason, there is a need to investigate the stability of the slopes that will be opened to urban use.

The importance of extensometer station was very critical for over 20 mm displacement and covering the cracks by mixture of melted plastics and asphalt and facilitates the infiltration of rainwater into the channel reduces the landslide or rock flow. This drainage patterns prevents the masses creep.

The lack of vegetation in the study area prevents benefiting from these effects, which are positive in terms of stability, and hence, there is a reduction in the holding forces that keep the slopes in balance. For this reason, enrichment in terms of vegetation is an important landslide preventing parameter in the region. However, the effect of vegetation on stability will be minimal for sliding surfaces up to 30 m in depth. The weathering causes the rocks to change to a great extent, the bond between the grains to weaken and disappear completely. The rocks, which weaken as a result of weathering in the study area, are easily eroded and change the slope angles and slope altitudes. The weathering observed in the rocks in the study area also contributes negatively to the stability problems. As a result of geotechnical analysis carried out in the study area, it is concluded that very large landslides will not be expected in the future. However, this result does not eliminate the possibility of landslide danger for urban areas and urban development areas. For this reason, especially in large-scale plans such as master zoning plan and application zoning plan, geological hazard in the region should be evaluated and landslide prevention methods appropriate for the site should be determined. In addition, research and development methods to prevent instability in the region are of special importance since the study area will be opened to urban use within the scope of the project.

As a result of laboratory experiments carried out on the soil samples, it was determined that the slope material was permeable, the cohesion value ranged between 1300 and 3800 kPa, and the internal friction angle ranged between 37.5 and 22.1, and according to the combined ground classification, the slope material consisted of Shale, low marly shale, and cracked limestone low RQD group rocks. From the stability analysis performed in the light of this information, it was concluded that the slopes S1, S2, and N1 were unstable, and the slopes S3 and N2 N3 were stable.

As a result of the field investigations, it was determined that the increase in the flow rate at the peak with the melting of snow in May and June caused severe erosion on the heel of the slopes and thus had a negative effect on the stability of the slope. In order to prevent this situation, which is effective on slopes N2 and N3, it is necessary to accumulate rock material in sizes that the stream cannot bear, or the stream should be improved. Sirnak City and the surrounding area, according to Turkey Earthquake Zone Map, are located in the first degree in the danger zone. Since these regions are within the influenced area of the South East Anatolian Fault, frequent earthquakes occur in the region and some tectonic movements occur due to these earthquakes. As a result of these movements, the stability of the slopes is compromised. For this reason, there is a need to investigate the stability of the slopes that will be opened to urban use.

Plants facilitate the infiltration of rainwater into the mass and slow down and reduce superficial flow. This prevents the masses from erosion. The roots of the plants, whose roots reach deep, are mechanical.

**5. Conclusions**

*Weight chart calculations for S13.*

**Table 10.**

**128**

**Chart Block height**

*Weight chart calculations for S12.*

**Table 9.**

**Chart Block height**

> **Block width**

> **Block width**

*Mining Techniques - Past, Present and Future*

**Block weight (ton)**

10 7 3 19,46 190,9026 11,40,951 500 Total 6317,967 342,9576 1418

**Block weight (ton)**

1 5 6 27,8 272,718 15,65,644 2 7 7 45,40,667 445,4394 24,62,219 3 12 8 88,96 872,6976 46,80,062 4 17 9 141,78 1390,862 73,69,786 5 18 7 116,76 1145,416 60,95,706 6 17 5 78,76,667 772,701 41,60,992 7 15 4 55,6 545,436 29,81,289 8 12 4 44,48 436,3488 24,15,031 9 9 3 25,02 245,4462 14,2408

1 9 9 75,06 736,3386 30,29,683 2 11 9 91,74 899,9694 40,41,929 3 15 9 125,1 1227,231 54,57,176 4 17 9 141,78 1390,862 61,64,799 5 18 7 116,76 1145,416 51,03364 6 17 5 78,76,667 772,701 34,91,555 7 14 4 51,89,333 509,0736 23,51,495 8 11 4 40,77,333 399,9864 18,79,746 9 9 3 25,02 245,4462 12,11,435

**Block weight (kN)**

**Block weight (kN)**

> **Block chart share (MPa)**

> **Block chart share (MPa)**

> > **Safety**

**Safety**

As a result of the laboratory experiments carried out on the soil samples, it was

10 7 3 19,46 190,9026 9,755,607 500 Total 7517,926 337,0674 1171

determined that the slope material was permeable, the cohesion value ranged between 0.1 and 0.38, the internal friction angle ranged between 17.5 and 22.4, and according to the combined ground classification, the slope material consisted of shale group plastic steep slopes in the top area of low RQD. From the stability analysis performed in the light of this information, it was concluded that the slopes in S1, S2, and N1 are unstable, and the slope S3 and N2 in open pit were stable. As a result of the field studies, it was determined that the increase in the flow rate at the peak with the melting of snow, especially in May and June, caused severe erosion on the heel of the slopes and thus had a negative effect on the stability of the slope. Chart patterns were obtained on the slopes S1, S2, and S3 to reduce sliding risk.

#### **Symbols**


**References**

1981

472 p

Sons; 1980

1974. pp. 1212-1217

2000. pp. 71-80

pp. 183-202

**31**:1627-1632

**131**

[1] Höek E, Bray J. Rock Slope Engineering. 3rd ed. Institution of Mining and Metallurgy: London, UK;

[2] Goodman RE. Methods of Geological Engineering in Discontinuous Rocks. St. Paul, MN: West Publishing Co.; 1976.

*DOI: http://dx.doi.org/10.5772/intechopen.91987*

Construction and Building Materials.

[10] Dramis F, Sorriso-Valvo M. Deep-Seated gravitational slope deformations,

related landslides and tectonics. Engineering Geology. 1994;**38**:231-243

[11] Ulusoy R. Uygulamalı Jeoteknik Bilgiler. Ankara: JMO Yayınları; 1989

[12] Gündüz L. The effects of pumice aggregate/cementations on the lowstrength concrete properties.

Construction and Building Materials.

Influence of a new type of additive on the performance of polymer- lightweight mortar composites. Cement and Concrete

[14] Gündüz L, Uğur İ. The effects of different fine and coarse pumice

[15] Gündüz L, Sarıısık A, Tozaçan B, Davraz M, Uğur İ, Çankıran O. Pumice Technology. Vol. 1. Isparta: Süleyman Demirel University; 1998. pp. 275-285

[16] Tosun Yİ, Cevizci H, Ceylan H. Landfill design for reclamation of Şırnak coal mine dumps—Shalefill stability and risk assessment. In: ICMEMT 2014, 11-12 July 2014, Prague, Czechoslovakia. 2014

[17] Tosun Yİ. A case study on use of foam concrete landfill on landslide hazardous area in Şırnak city province. In: XX Congress of the Carpathian Balkan Geological Association, Tirana, Albania, 24-26 September 2014. 2014

[18] Tosun Yİ. Shale stone and fly ash landfill use in land-slide hazardous area

aggregate/ cementations on the structural concrete properties without using any admixtures. Cement and Concrete Research. 2005;**35**(9):1859-1864

[13] Gündüz L, Bekar M, Şapcı N.

Composites. 2007;**29**:594-602

2008;**22**:135-142

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture…*

2008;**22**(5):721-728

[3] Goodman RE. Introduction to Rock Mechanics. New York: John Wiley and

[4] Pells PJN. The behaviour of fully bonded rock bolts. In: Advances in Rock Mechanics. Vol. 2: Part B. Washington, DC: National Academy of Sciences;

[5] Sjöberg J. Failure mechanism for high slopes in hard rock. In: Slope Stability in Surface Mining. Littleton, CO: Society of Mining, Metallurgy and Exploration;

[6] Sjöberg J, Sharp JC, Malorey DJ. Slope stability at Aznalcóllar. In:

[7] Chen B, Liu J. Experimental application of mineral admixtures in lightweight concrete with high strength and workability. Construction and Building Materials. 2008;**22**:s.655-s.659

Hustralid WA, MJ MC, DJA VZ, editors. Slope Stability in Surface Mining. Littleton, CO: Society for Mining, Metallurgy and Exploration; 2001.

[8] Demirboğa R, Orung I, Gül R. Effects of expanded perlite aggregate and mineral admixtures on the compressive strength of low-density concretes. Cement and Concrete Research. 2001;

[9] Demirdag S, Gündüz L. Strength properties of volcanic slag aggregate lightweight concrete for high performance masonry units.

### **Author details**

Yildırım İsmail Tosun Engineering Faculty, Mining Engineering Department, Şırnak University, Şırnak, Turkey

\*Address all correspondence to: yildirimismailtosun@gmail.com

© 2020 The Author(s). Licensee IntechOpen. This chapter is distributed under the terms of the Creative Commons Attribution License (http://creativecommons.org/licenses/ by/3.0), which permits unrestricted use, distribution, and reproduction in any medium, provided the original work is properly cited.

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture… DOI: http://dx.doi.org/10.5772/intechopen.91987*

#### **References**

**Symbols**

**Author details**

Turkey

**130**

Yildırım İsmail Tosun

Φ0

c<sup>0</sup> kg/cm<sup>2</sup> effective cohesion

*Mining Techniques - Past, Present and Future*

Φο internal friction angle

ο effective internal friction angle

γkmax g/cm<sup>3</sup> maximum dry unit volume weight

number

S1, S2, S3, S4, C1, C2 south and north landslide risk slopes No. 1, 2, 3, 4 S11, C11 sample taken from south and north landslide risk slopes

Engineering Faculty, Mining Engineering Department, Şırnak University, Şırnak,

© 2020 The Author(s). Licensee IntechOpen. This chapter is distributed under the terms of the Creative Commons Attribution License (http://creativecommons.org/licenses/ by/3.0), which permits unrestricted use, distribution, and reproduction in any medium,

\*Address all correspondence to: yildirimismailtosun@gmail.com

provided the original work is properly cited.

γs g/cm<sup>3</sup> grain unit volume weight k permeability coefficient

c kg/cm<sup>2</sup> cohesion

τ kg/cm<sup>2</sup> shear stress σ kg/cm<sup>2</sup> normal stress Is point load index Bs bending strength Ps compression strength Wopt optimum water content γNatural g/cm<sup>3</sup> natural unit volume weight γSaturated g/cm<sup>3</sup> saturated unit volume weight γDry g/cm<sup>3</sup> dry unit volume weight

[1] Höek E, Bray J. Rock Slope Engineering. 3rd ed. Institution of Mining and Metallurgy: London, UK; 1981

[2] Goodman RE. Methods of Geological Engineering in Discontinuous Rocks. St. Paul, MN: West Publishing Co.; 1976. 472 p

[3] Goodman RE. Introduction to Rock Mechanics. New York: John Wiley and Sons; 1980

[4] Pells PJN. The behaviour of fully bonded rock bolts. In: Advances in Rock Mechanics. Vol. 2: Part B. Washington, DC: National Academy of Sciences; 1974. pp. 1212-1217

[5] Sjöberg J. Failure mechanism for high slopes in hard rock. In: Slope Stability in Surface Mining. Littleton, CO: Society of Mining, Metallurgy and Exploration; 2000. pp. 71-80

[6] Sjöberg J, Sharp JC, Malorey DJ. Slope stability at Aznalcóllar. In: Hustralid WA, MJ MC, DJA VZ, editors. Slope Stability in Surface Mining. Littleton, CO: Society for Mining, Metallurgy and Exploration; 2001. pp. 183-202

[7] Chen B, Liu J. Experimental application of mineral admixtures in lightweight concrete with high strength and workability. Construction and Building Materials. 2008;**22**:s.655-s.659

[8] Demirboğa R, Orung I, Gül R. Effects of expanded perlite aggregate and mineral admixtures on the compressive strength of low-density concretes. Cement and Concrete Research. 2001; **31**:1627-1632

[9] Demirdag S, Gündüz L. Strength properties of volcanic slag aggregate lightweight concrete for high performance masonry units.

Construction and Building Materials. 2008;**22**:135-142

[10] Dramis F, Sorriso-Valvo M. Deep-Seated gravitational slope deformations, related landslides and tectonics. Engineering Geology. 1994;**38**:231-243

[11] Ulusoy R. Uygulamalı Jeoteknik Bilgiler. Ankara: JMO Yayınları; 1989

[12] Gündüz L. The effects of pumice aggregate/cementations on the lowstrength concrete properties. Construction and Building Materials. 2008;**22**(5):721-728

[13] Gündüz L, Bekar M, Şapcı N. Influence of a new type of additive on the performance of polymer- lightweight mortar composites. Cement and Concrete Composites. 2007;**29**:594-602

[14] Gündüz L, Uğur İ. The effects of different fine and coarse pumice aggregate/ cementations on the structural concrete properties without using any admixtures. Cement and Concrete Research. 2005;**35**(9):1859-1864

[15] Gündüz L, Sarıısık A, Tozaçan B, Davraz M, Uğur İ, Çankıran O. Pumice Technology. Vol. 1. Isparta: Süleyman Demirel University; 1998. pp. 275-285

[16] Tosun Yİ, Cevizci H, Ceylan H. Landfill design for reclamation of Şırnak coal mine dumps—Shalefill stability and risk assessment. In: ICMEMT 2014, 11-12 July 2014, Prague, Czechoslovakia. 2014

[17] Tosun Yİ. A case study on use of foam concrete landfill on landslide hazardous area in Şırnak city province. In: XX Congress of the Carpathian Balkan Geological Association, Tirana, Albania, 24-26 September 2014. 2014

[18] Tosun Yİ. Shale stone and fly ash landfill use in land-slide hazardous area in Sirnak city with foam concrete. GM Geomaterials Journal. 2014;**4**(4):141- 150. DOI: 10.4236/gm.2014.44014

[19] Tosun YI. Kalker, Marn ve Şeylin Sünme Karakterizasyonu—Bitümlü Gözenekli Agrega için Don— Mikrodalga Kurutma-Bilya Darbe Dayanım Testi ile Sünme Etüdü, AGGRE 2016. In: 8th Internatıonal Aggregates Symposıum, October 5-7, 2016, Istanbul, Turkey. 2016

[20] Rocscience. SWEDGE— Probabilistic Analysis of the Geometry and Stability of Surface Wedges. Toronto, Canada: Rocscience Ltd.; 2001 Available from: www.rocscience.com/

[21] Rocscience Ltd. ROCLAB Software for Calculating Hoek–Brown Rock Mass Strength. Toronto, Ontario: Rocscience Ltd; 2002. Available from: www. rocscience.com/

[22] Rocscience Ltd. SLIDE—2D Slope Stability Analysis for Rock and Soil Slopes. Toronto, Ontario: Rocscience Ltd; 2002. Available from: www. rocscience.com/

[23] Pritchard MA, Savigny KW. Numerical modelling of toppling. Canadian Geotechnical Journal. 1990;**27**: 823-834

[24] Pritchard MA, Savigny KW. The Heather Hill landslide: an example of a large scale toppling failure in a natural slope. Canadian Geotechnical Journal. 1991;**28**:410-422

[25] Sonmez H, Ulusay R. Modifications to the geological strength index (GSI) and their applicability to the stability of slopes. International Journal of Rock Mechanics and Mining Sciences. 1999; **36**(6):743-760

[26] Sageseta C, Sánchez JM, Cañizal J. A general solution for the required anchor force in rock slopes with toppling failure. International Journal of Rock

Mechanics and Mining Sciences. 2001; **38**:421-435

[37] Gündüz L. Use of quartet blends containing fly ash, scoria, perlitic pumice and cement to produce cellular hollow lightweight masonry blocks form on-load bearing walls. Construction and Building Materials. 2008;**22**:747-754

*DOI: http://dx.doi.org/10.5772/intechopen.91987*

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture…*

[38] Ulusoy R. Şev Stabilite Analizlerinde

[39] Görög P, Török Á. Slope stability assessment of weathered clay by using field data and computer modelling: a case study from Budapest. Natural Hazards and Earth System Sciences. 2007;**7**:417-422. Available from: www.nat-hazards-earth-syst-sci.net

Kullanılan Pratik Yöntemler ve Geoteknik Çalışmalar. MTA Yayınları,

[40] Görög P, Török Á. Stability Problems of Abandoned Clay Pits in Budapest, IAEG2006 Paper Number 295. London: The Geological Society of

London; 2006

**133**

Eğitim Serisi; 1982. p. 25

[27] Anbalagan R. Landslide hazard evaluation and zonation mapping in mountainous terrain. Engineering Geology. 1992;**32**:269-277

[28] Anonymous, a. GEO5—Engineering Manuals—Part 1, Part 2. 2011. Available from: http://www.finesoftware.eu/ geotechnical-software/

[29] Anonymous, b. GEO5—FEM— Theoretical Guide. 2011. Available from: http://www.finesoftware.eu/ geotechnical-software/

[30] Anonymous, c. Türkiye Deprem Bölgeleri Haritası. Ankara: Afet ve Acil durum Yönetimi Başkanlığı Deprem Dairesi Başkanlığı; 2012

[31] Duncan JM. Landslides: Investigation and mitigation - Chapter 13. In: Soil Slope Stability Analysis. Washington, DC: Transportation Research Board; 1996. pp. 337-371. Special Report 247

[32] ASTM. Standard Test Method for Direct Shear Test of Soils under Consolidated Drained Condition1990. pp. D3080-D3090

[33] Dershowitz WS, Einstein HH. Characterizing rock joint geometry with joint system models. Rock Mechanics and Rock Engineering. 1988;**20**(1):21-51

[34] Höek E. Estimating the stability of excavated slopes in opencast mines. Institution of Mining and Metallurgy. 1970;**A105**:A132

[35] Fell R. Landslide risk assessment and acceptable risk. Canadian Geotechnical Journal. 1994;**31**:261-272

[36] Erdoğan TY. Beton. Ankara: ODTÜ Geliştirme Vakfı Yayıncılık ve İletişim A.Ş; 2003

*Anchorage Pile Strengthening of Shale Slopes and Cementing Falling Stone Blocks by Mixture… DOI: http://dx.doi.org/10.5772/intechopen.91987*

[37] Gündüz L. Use of quartet blends containing fly ash, scoria, perlitic pumice and cement to produce cellular hollow lightweight masonry blocks form on-load bearing walls. Construction and Building Materials. 2008;**22**:747-754

in Sirnak city with foam concrete. GM Geomaterials Journal. 2014;**4**(4):141- 150. DOI: 10.4236/gm.2014.44014

*Mining Techniques - Past, Present and Future*

Mechanics and Mining Sciences. 2001;

[28] Anonymous, a. GEO5—Engineering Manuals—Part 1, Part 2. 2011. Available from: http://www.finesoftware.eu/

[29] Anonymous, b. GEO5—FEM— Theoretical Guide. 2011. Available from:

[30] Anonymous, c. Türkiye Deprem Bölgeleri Haritası. Ankara: Afet ve Acil durum Yönetimi Başkanlığı Deprem

Investigation and mitigation - Chapter 13. In: Soil Slope Stability Analysis. Washington, DC: Transportation Research Board; 1996. pp. 337-371.

[32] ASTM. Standard Test Method for Direct Shear Test of Soils under Consolidated Drained Condition1990.

[33] Dershowitz WS, Einstein HH. Characterizing rock joint geometry with joint system models. Rock Mechanics and Rock Engineering. 1988;**20**(1):21-51

[34] Höek E. Estimating the stability of excavated slopes in opencast mines. Institution of Mining and Metallurgy.

[35] Fell R. Landslide risk assessment and acceptable risk. Canadian

Geotechnical Journal. 1994;**31**:261-272

[36] Erdoğan TY. Beton. Ankara: ODTÜ Geliştirme Vakfı Yayıncılık ve İletişim

http://www.finesoftware.eu/ geotechnical-software/

Dairesi Başkanlığı; 2012

Special Report 247

pp. D3080-D3090

1970;**A105**:A132

A.Ş; 2003

[31] Duncan JM. Landslides:

[27] Anbalagan R. Landslide hazard evaluation and zonation mapping in mountainous terrain. Engineering

Geology. 1992;**32**:269-277

geotechnical-software/

**38**:421-435

[19] Tosun YI. Kalker, Marn ve Şeylin Sünme Karakterizasyonu—Bitümlü Gözenekli Agrega için Don— Mikrodalga Kurutma-Bilya Darbe Dayanım Testi ile Sünme Etüdü, AGGRE 2016. In: 8th Internatıonal Aggregates Symposıum, October 5-7,

2016, Istanbul, Turkey. 2016

[20] Rocscience. SWEDGE—

rocscience.com/

rocscience.com/

1991;**28**:410-422

**36**(6):743-760

**132**

823-834

Probabilistic Analysis of the Geometry and Stability of Surface Wedges. Toronto, Canada: Rocscience Ltd.; 2001 Available from: www.rocscience.com/

[21] Rocscience Ltd. ROCLAB Software for Calculating Hoek–Brown Rock Mass Strength. Toronto, Ontario: Rocscience Ltd; 2002. Available from: www.

[22] Rocscience Ltd. SLIDE—2D Slope Stability Analysis for Rock and Soil Slopes. Toronto, Ontario: Rocscience Ltd; 2002. Available from: www.

[23] Pritchard MA, Savigny KW. Numerical modelling of toppling. Canadian Geotechnical Journal. 1990;**27**:

[24] Pritchard MA, Savigny KW. The Heather Hill landslide: an example of a large scale toppling failure in a natural slope. Canadian Geotechnical Journal.

[25] Sonmez H, Ulusay R. Modifications to the geological strength index (GSI) and their applicability to the stability of slopes. International Journal of Rock Mechanics and Mining Sciences. 1999;

[26] Sageseta C, Sánchez JM, Cañizal J. A general solution for the required anchor force in rock slopes with toppling failure. International Journal of Rock

[38] Ulusoy R. Şev Stabilite Analizlerinde Kullanılan Pratik Yöntemler ve Geoteknik Çalışmalar. MTA Yayınları, Eğitim Serisi; 1982. p. 25

[39] Görög P, Török Á. Slope stability assessment of weathered clay by using field data and computer modelling: a case study from Budapest. Natural Hazards and Earth System Sciences. 2007;**7**:417-422. Available from: www.nat-hazards-earth-syst-sci.net

[40] Görög P, Török Á. Stability Problems of Abandoned Clay Pits in Budapest, IAEG2006 Paper Number 295. London: The Geological Society of London; 2006

**Chapter 8**

East of Slovakia

realization of the reclaiming work.

model, test statistics, reclaiming

**1. Introduction**

**135**

*Vladimír Sedlák*

**Abstract**

Specific Solution of Deformation

Applications to Reclaiming the

Vector in Land Subsidence for GIS

Abandoned Magnesite Mine in the

Mining activity influences on the environment belong to the most negative industrial influences. Mining subsidence on the earth surface is a result of underground mining. The present study deals with the theory of specific procedures for solving the deformation vector in the case of an objective disturbance of data homogeneity in the geodetic network structure of the monitoring station in monitoring mining subsidence. The theory was developed for the mining subsidence created on the earth surface of the mining landscape, where the abandoned magnesite mine Košice-Bankov in the East of Slovakia was operated for many decades in the twentieth century. The achieved results and outputs were implemented into the GIS tools for the plan of the process of gradual reclaiming the entire mining landscape of Košice-Bankov. The aim of the deformation measurements was to determine the exact boundaries of the subsidence edges with the residual movement zones for the purpose of comprehensive reclaiming the devastated mining landscape. Some numerical and graphical results from the deformation vectors survey in the abandoned magnesite mine Košice-Bankov are presented. The obtained results in GIS were supplied for the needs of the Municipality of the city of Košice to the

**Keywords:** subsidence, deformation vector, geodetic network, Gauss-Markov

At present, with an extremely sharp increase in people's material needs, priority must be given to their security from any economic prosperity of many countries around the world. As the extraction and processing mineral resources increase, so does the economic prosperity of the country. To protect the environment, which should be an intact ecosystem, it is necessary to protect the lives of people and their property from adverse industrial impacts. One of the most negative industrial impacts on the whole ecosystem is the adverse impact of any mining activity. The land subsidence (mining subsidence, hereinafter referred to as subsidence)

#### **Chapter 8**

## Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming the Abandoned Magnesite Mine in the East of Slovakia

*Vladimír Sedlák*

### **Abstract**

Mining activity influences on the environment belong to the most negative industrial influences. Mining subsidence on the earth surface is a result of underground mining. The present study deals with the theory of specific procedures for solving the deformation vector in the case of an objective disturbance of data homogeneity in the geodetic network structure of the monitoring station in monitoring mining subsidence. The theory was developed for the mining subsidence created on the earth surface of the mining landscape, where the abandoned magnesite mine Košice-Bankov in the East of Slovakia was operated for many decades in the twentieth century. The achieved results and outputs were implemented into the GIS tools for the plan of the process of gradual reclaiming the entire mining landscape of Košice-Bankov. The aim of the deformation measurements was to determine the exact boundaries of the subsidence edges with the residual movement zones for the purpose of comprehensive reclaiming the devastated mining landscape. Some numerical and graphical results from the deformation vectors survey in the abandoned magnesite mine Košice-Bankov are presented. The obtained results in GIS were supplied for the needs of the Municipality of the city of Košice to the realization of the reclaiming work.

**Keywords:** subsidence, deformation vector, geodetic network, Gauss-Markov model, test statistics, reclaiming

#### **1. Introduction**

At present, with an extremely sharp increase in people's material needs, priority must be given to their security from any economic prosperity of many countries around the world. As the extraction and processing mineral resources increase, so does the economic prosperity of the country. To protect the environment, which should be an intact ecosystem, it is necessary to protect the lives of people and their property from adverse industrial impacts. One of the most negative industrial impacts on the whole ecosystem is the adverse impact of any mining activity. The land subsidence (mining subsidence, hereinafter referred to as subsidence)

is created on the earth surface as a result of underground extraction of the mineral deposits especially at the chamber mining by a caving method [1, 2]. In many cases, the mine subsidence represents the large-scale and very deep downthrown blocks that is dangerous for any movement of people in them [1–7]. As at the deep mining there are many large voids created in the rock massif (especially during the aforementioned mining by the chamber method), their collapse occurs with the manifestations of deformations on the earth surface mostly in the form of subsidence. The collapse of these cavities can occur at different time horizons, i.e., from the commencement of mining up to several years, or even decades or more after the end of mining. Especially for endangering the lives of people and their property, the most dangerous are the sudden and unexpected formation of invasions of the earth surface over the mined rock massif, which often happens even in some abandoned mines [6, 8, 9].

surveying technology, are increasingly being applied to deformation measurements [14, 15, 18–26]. The deformation vectors are the outputs of 3D data processing from the abovementioned geodetic measurements of deformations (deformation measurements) of the earth surface or building structure objects, but several geodetic measurements are combined and thus confirmed by some physical measurements based on the measurement of stress and strain states. By their size and position in 3D space, deformation vectors can provide a global overview not only of the nature of the current state of deformation of the monitoring subsidence in the mining area, but also of the nature of the further development of these deformations with the required time predictions [1, 2, 10, 19]. At the same time, when monitoring particularly important buildings located close to the subsidence, additional geophysical measurements of stress–strain states of the rock massif must be carried out underground (in the extracted mining space). Achievement of the unstressing states in the rock massif above the extracted space can be realized in many cases by the

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming…*

Periodically repeating measurements on parts respectively on the whole geodetic network of some monitoring stations to measure deformations of the earth surface (e.g., subsidence) or construction objects and engineering structures can sometimes be complicated for objective reasons. Most of the monitoring stations consist of a network of firmly stabilized points (geodetic network) on the surveyed earth surface or on surveyed buildings and other engineering structures. By various geodetic methods, measurements are made on these points of the geodetic network to determine their 3D positions (i.e., coordinates *X*, *Y*, and *Z*) in the given Cartesian coordinate system [5, 14]. During the realization of long-term (multiyear or even several decades) geodetic measurements, undesirable and unpredictable obstructions and interventions into the monitoring station and so into the geometric structure of the geodetic network may also occur. In many cases, such undesirable interventions into the structure of the geodetic network include predominantly loss or damage to points due to uncontrolled earthworks or construction work in the territory of the monitoring station. Due to such undesirable interference with the monitoring station, the geometric and data structure of the points of the geodetic network is not homogeneous during all periodic deformation measurements.

This means that the measured elements of the applied geodetic measurements at

In the evaluation of deformation vectors and their 3D modelling, the time factor of the gradual creation of the subsidence over the mined-out space underground plays an important role in its overall evaluation of the earth ground movements. The possibility in improving polynomial modeling of the subsidence is conditioned by the knowledge to detect position of the so-called breakpoints, i.e., the points in the surface in which the subsidence border with a zone of breaches and bursts start to develop above the mined mineral deposit. This means that the

the points of the geodetic network of the monitoring station at individual time epochs are no longer identical and cannot be maintained as they were at the time of the initial (primary, i.e., zero) measured epoch. When changing the geometrical structure of the points of the geodetic network, it is no longer possible to maintain, in particular, the spatial configuration of the measuring sights between the determined points of the network. Disturbance of stability or destruction and the loss of multiple points of the geodetic network also results in a disturbance of the geometric and thus data structure of the whole network of the monitoring station. In such cases, the re-stabilization of destroyed or lost points (in particular object points) at the original or other places of the geodetic network will not help either. Nor will it help to replace the original measured variables (which are no longer measurable in the subsequent monitoring epochs due to damage of the geodetic network points)

so-called destress blasting [27, 28].

*DOI: http://dx.doi.org/10.5772/intechopen.91461*

by other measured variables [1, 10, 25, 29].

**137**

In the sense of several scientific studies, as well as theoretical and practical knowledge on mining impacts on the earth surface, it is very difficult to accurately predict how long the movements in the subsidence will be in existence and when these movements will be ended [1, 2, 10–12]. Current statistics resulting from many deformation measurements in the subsidence confirm that around 60–90% of total subsidence movements occur within the first few weeks up to months after the excavation of the underground space in the rock massif. The remaining movements in the subsidence persist even after the end of mining, and at a decreasing rate these movements can last for 3–5 years and longer and in rare cases even several decades.

In order to protect the lives of people and their property, as well as the complex environment, it is essential to constantly investigate any movements in the subsidence on the earth surface even after mining has ended [2, 13, 14]. The most investigated movements in the subsidence with their prediction based on threedimensional (3D) modeling are on coal fields [3, 5, 11, 12, 14–17].

The nature and magnitude of the subsidence on the earth surface depends mainly on tectonic and geological conditions and also on the overload of the rock massif above the excavated space. Knowing the range, i.e., localization of the edges (boundaries) of the subsidence in mining territories, may provide for more precise placement of technical barriers (fencings and warning boards) and thus help to prevent persons and animals from entering these danger zones. Geodetic methods for investigation of deformation vectors, which can be derived from the processing of some specific geodetic measurements at the monitoring stations based on these mining territories, are the priority methods for determining the extent of movement in the subsidence on the earth surface above the excavated underground space. The deformation vectors in their 3D model concept most markedly characterize any movements of the earth surface, buildings, and other civil engineering structures located in the mining territory with the occurrence of the subsidence. In many cases, 3D modeling of the deformation vectors is based on regular (periodic) monitoring of the spatial changes in suitably structured points of the geodetic network of a monitoring station located on the earth surface or on buildings and civil engineering structures. Such periodic monitoring of the spatial changes of 3D coordinates of the geodetic network points is mostly realized by the application of classical geodetic terrestrial methods (tacheometry, trigonometry, traverse surveying, leveling, etc.), which are used extensively up to now or are currently increasingly using advanced methods of the satellite technologies within the Global Navigation Satellite Systems (GNSS) based on the US Global Positioning System (GPS), Russian Global Navigation Sputnik System (GLONAS), European Galileo, and Chinese Compass (BeiDou II). At present, very specific measurement technologies such as Interferometric Synthetic Aperture Radar (InSAR), i.e., the radar surveying technology, and Light Detection and Ranging (LiDAR), i.e., the laser

#### *Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming… DOI: http://dx.doi.org/10.5772/intechopen.91461*

surveying technology, are increasingly being applied to deformation measurements [14, 15, 18–26]. The deformation vectors are the outputs of 3D data processing from the abovementioned geodetic measurements of deformations (deformation measurements) of the earth surface or building structure objects, but several geodetic measurements are combined and thus confirmed by some physical measurements based on the measurement of stress and strain states. By their size and position in 3D space, deformation vectors can provide a global overview not only of the nature of the current state of deformation of the monitoring subsidence in the mining area, but also of the nature of the further development of these deformations with the required time predictions [1, 2, 10, 19]. At the same time, when monitoring particularly important buildings located close to the subsidence, additional geophysical measurements of stress–strain states of the rock massif must be carried out underground (in the extracted mining space). Achievement of the unstressing states in the rock massif above the extracted space can be realized in many cases by the so-called destress blasting [27, 28].

Periodically repeating measurements on parts respectively on the whole geodetic network of some monitoring stations to measure deformations of the earth surface (e.g., subsidence) or construction objects and engineering structures can sometimes be complicated for objective reasons. Most of the monitoring stations consist of a network of firmly stabilized points (geodetic network) on the surveyed earth surface or on surveyed buildings and other engineering structures. By various geodetic methods, measurements are made on these points of the geodetic network to determine their 3D positions (i.e., coordinates *X*, *Y*, and *Z*) in the given Cartesian coordinate system [5, 14]. During the realization of long-term (multiyear or even several decades) geodetic measurements, undesirable and unpredictable obstructions and interventions into the monitoring station and so into the geometric structure of the geodetic network may also occur. In many cases, such undesirable interventions into the structure of the geodetic network include predominantly loss or damage to points due to uncontrolled earthworks or construction work in the territory of the monitoring station. Due to such undesirable interference with the monitoring station, the geometric and data structure of the points of the geodetic network is not homogeneous during all periodic deformation measurements.

This means that the measured elements of the applied geodetic measurements at the points of the geodetic network of the monitoring station at individual time epochs are no longer identical and cannot be maintained as they were at the time of the initial (primary, i.e., zero) measured epoch. When changing the geometrical structure of the points of the geodetic network, it is no longer possible to maintain, in particular, the spatial configuration of the measuring sights between the determined points of the network. Disturbance of stability or destruction and the loss of multiple points of the geodetic network also results in a disturbance of the geometric and thus data structure of the whole network of the monitoring station. In such cases, the re-stabilization of destroyed or lost points (in particular object points) at the original or other places of the geodetic network will not help either. Nor will it help to replace the original measured variables (which are no longer measurable in the subsequent monitoring epochs due to damage of the geodetic network points) by other measured variables [1, 10, 25, 29].

In the evaluation of deformation vectors and their 3D modelling, the time factor of the gradual creation of the subsidence over the mined-out space underground plays an important role in its overall evaluation of the earth ground movements.

The possibility in improving polynomial modeling of the subsidence is conditioned by the knowledge to detect position of the so-called breakpoints, i.e., the points in the surface in which the subsidence border with a zone of breaches and bursts start to develop above the mined mineral deposit. This means that the

is created on the earth surface as a result of underground extraction of the mineral deposits especially at the chamber mining by a caving method [1, 2]. In many cases, the mine subsidence represents the large-scale and very deep downthrown blocks that is dangerous for any movement of people in them [1–7]. As at the deep mining there are many large voids created in the rock massif (especially during the aforementioned mining by the chamber method), their collapse occurs with the manifestations of deformations on the earth surface mostly in the form of subsidence. The collapse of these cavities can occur at different time horizons, i.e., from the commencement of mining up to several years, or even decades or more after the end of mining. Especially for endangering the lives of people and their property, the most dangerous are the sudden and unexpected formation of invasions of the earth surface over the mined rock massif, which often happens even in some abandoned

In the sense of several scientific studies, as well as theoretical and practical knowledge on mining impacts on the earth surface, it is very difficult to accurately predict how long the movements in the subsidence will be in existence and when these movements will be ended [1, 2, 10–12]. Current statistics resulting from many deformation measurements in the subsidence confirm that around 60–90% of total subsidence movements occur within the first few weeks up to months after the excavation of the underground space in the rock massif. The remaining movements in the subsidence persist even after the end of mining, and at a decreasing rate these movements can last for 3–5 years and longer and in rare cases even several decades. In order to protect the lives of people and their property, as well as the complex environment, it is essential to constantly investigate any movements in the subsidence on the earth surface even after mining has ended [2, 13, 14]. The most investigated movements in the subsidence with their prediction based on three-

The nature and magnitude of the subsidence on the earth surface depends mainly on tectonic and geological conditions and also on the overload of the rock massif above the excavated space. Knowing the range, i.e., localization of the edges (boundaries) of the subsidence in mining territories, may provide for more precise placement of technical barriers (fencings and warning boards) and thus help to prevent persons and animals from entering these danger zones. Geodetic methods for investigation of deformation vectors, which can be derived from the processing of some specific geodetic measurements at the monitoring stations based on these mining territories, are the priority methods for determining the extent of movement in the subsidence on the earth surface above the excavated underground space. The deformation vectors in their 3D model concept most markedly characterize any movements of the earth surface, buildings, and other civil engineering structures located in the mining territory with the occurrence of the subsidence. In many cases, 3D modeling of the deformation vectors is based on regular (periodic) monitoring of the spatial changes in suitably structured points of the geodetic network of a monitoring station located on the earth surface or on buildings and civil engineering structures. Such periodic monitoring of the spatial changes of 3D coordinates of the geodetic network points is mostly realized by the application of classical geodetic terrestrial methods (tacheometry, trigonometry, traverse surveying, leveling, etc.), which are used extensively up to now or are currently increasingly using advanced methods of the satellite technologies within the Global Navigation Satellite Systems (GNSS) based on the US Global Positioning System (GPS), Russian Global Navigation Sputnik System (GLONAS), European Galileo, and Chinese Compass (BeiDou II). At present, very specific measurement technologies such as Interferometric Synthetic Aperture Radar (InSAR), i.e., the radar surveying technology, and Light Detection and Ranging (LiDAR), i.e., the laser

dimensional (3D) modeling are on coal fields [3, 5, 11, 12, 14–17].

mines [6, 8, 9].

*Mining Techniques - Past, Present and Future*

**136**

breakpoints determine the edges of the subsurface at which the naturally consistent coherent of the earth surface is broken and the subsidence begins to form. 3D deformation vector models help to support the location of the breakpoints [10–13, 18, 21].

#### **2. Theory of the deformation vector specific solution**

As already mentioned in the Introduction, the geometric and thus data structure of the geodetic network of the monitoring station in the subsidence may be changed by some external intervention, such as some unforeseen earthworks and construction works at the monitoring station. Estimation of the structures of geodetic networks based on the Gauss-Markov model is the most used and the most effective method for their processing. In determining the statistical formulation of the Gauss-Markov model, we start from the following equations [11, 12, 17, 30–33]:

$$\boldsymbol{\sigma} = \mathbf{A} \left( \hat{\mathbf{C}} - \mathbf{C}^{\boldsymbol{\sigma}} \right) - \left( \mathbf{L}\_{(\mathbf{0})} - \mathbf{L}^{\boldsymbol{\sigma}} \right) = \mathbf{A} d \hat{\mathbf{C}} - d \mathbf{L} \tag{1}$$

$$
\Sigma\_L = \sigma\_0^2 \mathbf{Q}\_L \tag{2}
$$

*<sup>C</sup>*^ð Þ*<sup>i</sup>* <sup>¼</sup> *<sup>C</sup><sup>o</sup>* <sup>þ</sup> *<sup>A</sup>TQ*�**<sup>1</sup>**

*DOI: http://dx.doi.org/10.5772/intechopen.91461*

transformed into the following equations:

According to Eqs. (6)–(9), the vectors ^

*Q*�<sup>1</sup> *<sup>L</sup> A* � ��<sup>1</sup>

*Q*�<sup>1</sup> *<sup>L</sup> A* � ��<sup>1</sup>

<sup>þ</sup> *<sup>A</sup><sup>T</sup>*

<sup>þ</sup> *<sup>A</sup><sup>T</sup>*

epochs *t*ð Þ <sup>0</sup> and *t*ð Þ*<sup>i</sup>* .

^ *<sup>C</sup>*ð Þ <sup>0</sup> <sup>¼</sup> *<sup>C</sup><sup>o</sup>*

^ *<sup>C</sup>*ð Þ*<sup>i</sup>* <sup>¼</sup> *<sup>C</sup><sup>o</sup>*

the subsidence.

structures *<sup>d</sup>C*^ and *<sup>d</sup>*^

equation will be valid:

**139**

of the deformation vector *<sup>d</sup>*^

*d*^ *<sup>C</sup>* <sup>¼</sup> *<sup>C</sup><sup>o</sup>*

the form

*<sup>L</sup> <sup>A</sup>* � ��**<sup>1</sup>**

*ATQ*�**<sup>1</sup>**

• Thus, for the deformation vector *dC*^ will be valid in the following equation:

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming…*

where *L*ð Þ **<sup>0</sup>** and *L*ð Þ*<sup>i</sup>* are the vectors of the observed magnitude values in the

*<sup>t</sup>*ð Þ*<sup>i</sup>* is changed. Then the original matrixes and vectors *<sup>A</sup>*, *QL*, *<sup>C</sup><sup>o</sup>* and *Lo* are

Furthermore, we consider the case when there is a change in the geometric and thus in the data structure of the geodetic network of the monitoring station between the individual epochs of measurements. It means that the geometric and data structure of the geodetic network between the basic epoch *t*ð Þ <sup>0</sup> and the actual epoch

*<sup>C</sup>*ð Þ <sup>0</sup> and ^

*<sup>L</sup> <sup>L</sup>*ð Þ <sup>0</sup> � *<sup>L</sup>* � �*<sup>o</sup>* <sup>¼</sup> *<sup>C</sup><sup>o</sup>*

*<sup>L</sup> <sup>L</sup>*ð Þ*<sup>i</sup>* � *<sup>L</sup>* � �*<sup>o</sup>* <sup>¼</sup> *<sup>C</sup><sup>o</sup>*

*<sup>C</sup>*ð Þ*<sup>i</sup>* � ^

which not only expresses the 3D changes in the coordinates of the geodetic

obtained will not provide reliable data for testing the particular deformations in

The proposed and presented theory of the specific solution of the deformation vector in the case of any structural changes in the geodetic network will be acceptable for its proving in an analytical way, if we compare the deformation vector

<sup>þ</sup> *G L*ð Þ*<sup>i</sup>* � *<sup>L</sup>* � �*<sup>o</sup>* h i � *<sup>C</sup><sup>o</sup>* <sup>þ</sup> *G L*ð Þ <sup>0</sup> � *Lo* � � � �

*C* expressed according to Eqs. (5) and (12). Then the structure

<sup>¼</sup> *G L*ð Þ*<sup>i</sup>* � *Lo* � � � *G L*ð Þ <sup>0</sup> � *Lo* � � <sup>þ</sup> *<sup>C</sup><sup>o</sup>* � *<sup>C</sup><sup>o</sup>* (13)

*C* is expressed according to Eq. (12), and the further

coordinates of the observed points in the epochs *t*ð Þ <sup>0</sup> and *t*ð Þ*<sup>i</sup>* will be determined:

and then the deformation vector *dC*^ is expressed according to Eq. (5) in

*AT Q*�<sup>1</sup>

*AT Q*�<sup>1</sup>

> *d*^ *<sup>C</sup>* <sup>¼</sup> ^

network points between the individual epochs of measurement, but such deformation vector can also express changes in the overall structure (geometric

and data structure) of the geodetic network. The deformation vector *<sup>d</sup>*^

*<sup>L</sup> <sup>L</sup>*ð Þ*<sup>i</sup>* � *<sup>L</sup><sup>o</sup>* � � <sup>¼</sup> *<sup>C</sup><sup>o</sup>* <sup>þ</sup> *G L*ð Þ*<sup>i</sup>* � *<sup>L</sup><sup>o</sup>* � � (4)

*A* ¼ *A* þ *dA* (6) *QL* ¼ *QL* þ *dQL* (7) *<sup>C</sup><sup>o</sup>* <sup>¼</sup> *<sup>C</sup><sup>o</sup>* <sup>þ</sup> *<sup>d</sup>C<sup>o</sup>* (8) *<sup>L</sup><sup>o</sup>* <sup>¼</sup> *<sup>L</sup><sup>o</sup>* <sup>þ</sup> *<sup>d</sup><sup>L</sup>* (9)

*C*ð Þ*<sup>i</sup>* of the adjusted 3D

*C*ð Þ **<sup>0</sup>** (12)

<sup>þ</sup> *G L*ð Þ <sup>0</sup> � *<sup>L</sup>* � �*<sup>o</sup>* (10)

<sup>þ</sup> *G L*ð Þ*<sup>i</sup>* � *<sup>L</sup>*<sup>0</sup> � � (11)

*C* thus

*<sup>d</sup>C*^ <sup>¼</sup> *<sup>C</sup>*^ð Þ*<sup>i</sup>* � *<sup>C</sup>*^ð Þ **<sup>0</sup>** <sup>¼</sup> *G L*ð Þ*<sup>i</sup>* � *<sup>L</sup><sup>o</sup>* � � (5)

where *v* is the vector of corrections of the measured (observed) values *L*,*A* is the configuration (modeling) matrix of the geodetic network (otherwise called also the Jacobian matrix), i.e., the matrix of the partial derivatives of functions *Lo* <sup>¼</sup> *<sup>f</sup> <sup>C</sup><sup>o</sup>* ð Þ by the vector *C<sup>o</sup>* ,*C*^ is the vector of the aligned 3D coordinate values, *C<sup>o</sup>* is the vector of the approximate 3D coordinate values, *L*ð Þ <sup>0</sup> is the vector of the approximate observation magnitude values of the observed elements in the first measuring epoch *<sup>t</sup>*ð Þ <sup>0</sup> , *<sup>L</sup><sup>o</sup>* is the vector of the approximate observation magnitude values of the observed elements, *dC*^ is the deformation vector, *dL* is the vector of the measured values supplements, *Σ<sup>L</sup>* is the covariance matrix of the measured values, *σ*<sup>2</sup> <sup>0</sup> is a priori variance, and *QL* is the cofactor matrix of the observations.

It will also be appeared in the changed structures, let us say in a size of the matrixes and vectors *A*, *QL*, *C<sup>o</sup>* and *Lo* . These matrixes and vectors enter into the presupposed model of a network adjustment following out from the Gauss-Markov model [11, 12, 29–31].

#### **2.1 Deformation vector**

If between monitoring epochs there are no changes in the geometrical and observational structure of the geodetic network, then the matrixes and vectors *A*, *QL*, *C<sup>o</sup>* and *L<sup>o</sup>* remain identical for each epoch. Only in such case the deformation vector *dC*^ can be determined by a conventional procedure according to the following model [2, 16, 17]:

• In the basic (first) monitoring epoch *<sup>t</sup>*ð Þ <sup>0</sup> , we have the vector *<sup>C</sup>*^ð Þ <sup>0</sup> of the adjusted 3D coordinates of the observed points which are obtained according to the Gauss-Markov model:

$$\hat{\mathbf{C}}\_{(\mathfrak{0})} = \mathbf{C}^{\bullet} + \left(\mathbf{A}^{T}\mathbf{Q}\_{L}^{-1}\mathbf{A}\right)^{-1}\mathbf{A}^{T}\mathbf{Q}\_{L}^{-1}(L\_{(\mathfrak{0})} - L^{\bullet}) = \mathbf{C}^{\bullet} + \mathbf{G}\left(L\_{(\mathfrak{0})} - L^{\bullet}\right) \tag{3}$$

• In the other following epochs *<sup>t</sup>*ð Þ*<sup>i</sup>* , we also obtain the vector *<sup>C</sup>*^ð Þ*<sup>i</sup>* of the adjusted 3D coordinates of the observed points according to the equation:

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming… DOI: http://dx.doi.org/10.5772/intechopen.91461*

$$\hat{\mathbf{C}}\_{(i)} = \mathbf{C}^{o} + \left(\mathbf{A}^{T}\mathbf{Q}\_{L}^{-1}\mathbf{A}\right)^{-1}\mathbf{A}^{T}\mathbf{Q}\_{L}^{-1}(\mathbf{L}\_{(i)} - \mathbf{L}^{o}) = \mathbf{C}^{o} + \mathbf{G}\left(\mathbf{L}\_{(i)} - \mathbf{L}^{o}\right) \tag{4}$$

• Thus, for the deformation vector *dC*^ will be valid in the following equation:

$$d\hat{\mathbf{C}} = \hat{\mathbf{C}}\_{(\hat{\mathbf{t}})} - \hat{\mathbf{C}}\_{(\mathbf{0})} = \mathbf{G} \left( \mathbf{L}\_{(\hat{\mathbf{t}})} - \mathbf{L}^{\sigma} \right) \tag{5}$$

where *L*ð Þ **<sup>0</sup>** and *L*ð Þ*<sup>i</sup>* are the vectors of the observed magnitude values in the epochs *t*ð Þ <sup>0</sup> and *t*ð Þ*<sup>i</sup>* .

Furthermore, we consider the case when there is a change in the geometric and thus in the data structure of the geodetic network of the monitoring station between the individual epochs of measurements. It means that the geometric and data structure of the geodetic network between the basic epoch *t*ð Þ <sup>0</sup> and the actual epoch *<sup>t</sup>*ð Þ*<sup>i</sup>* is changed. Then the original matrixes and vectors *<sup>A</sup>*, *QL*, *<sup>C</sup><sup>o</sup>* and *Lo* are transformed into the following equations:

$$
\overline{\mathbf{A}} = \mathbf{A} + d\mathbf{A} \tag{6}
$$

$$\overline{\mathbf{Q}}\_{L} = \mathbf{Q}\_{L} + d\mathbf{Q}\_{L} \tag{7}$$

$$\overline{\mathbf{C}}^{\boldsymbol{\theta}} = \mathbf{C}^{\boldsymbol{\circ}} + d\mathbf{C}^{\boldsymbol{\circ}} \tag{8}$$

$$
\mathbf{L}^o = \mathbf{L}^o + d\mathbf{L} \tag{9}
$$

According to Eqs. (6)–(9), the vectors ^ *<sup>C</sup>*ð Þ <sup>0</sup> and ^ *C*ð Þ*<sup>i</sup>* of the adjusted 3D coordinates of the observed points in the epochs *t*ð Þ <sup>0</sup> and *t*ð Þ*<sup>i</sup>* will be determined:

$$\hat{\overline{\mathbf{C}}}\_{(0)} = \overline{\mathbf{C}}^{\rho} + \left(\overline{\mathbf{A}}^{T} \overline{\mathbf{Q}}\_{L}^{-1} \overline{\mathbf{A}}\right)^{-1} \overline{\mathbf{A}}^{T} \overline{\mathbf{Q}}\_{L}^{-1} (\mathbf{L}\_{(0)} - \overline{\mathbf{L}}^{\rho}) = \overline{\mathbf{C}}^{\rho} + \overline{\mathbf{G}} (\mathbf{L}\_{(0)} - \overline{\mathbf{L}}^{\rho}) \tag{10}$$

$$
\hat{\overline{\mathbf{C}}}\_{(i)} = \overline{\mathbf{C}}^{\boldsymbol{\rho}} + \left(\overline{\mathbf{A}}^{T} \overline{\mathbf{Q}}\_{L}^{-1} \overline{\mathbf{A}}\right)^{-1} \overline{\mathbf{A}}^{T} \overline{\mathbf{Q}}\_{L}^{-1} (\mathbf{L}\_{(i)} - \overline{\mathbf{L}}^{\boldsymbol{\rho}}) = \overline{\mathbf{C}}^{\boldsymbol{\rho}} + \overline{\mathbf{G}} \left(\mathbf{L}\_{(i)} - \overline{\mathbf{L}}^{\boldsymbol{\rho}}\right) \tag{11}
$$

and then the deformation vector *dC*^ is expressed according to Eq. (5) in the form

$$d\hat{\overline{\mathbf{C}}} = \hat{\overline{\mathbf{C}}}\_{(i)} - \hat{\overline{\mathbf{C}}}\_{(\mathbf{0})} \tag{12}$$

which not only expresses the 3D changes in the coordinates of the geodetic network points between the individual epochs of measurement, but such deformation vector can also express changes in the overall structure (geometric and data structure) of the geodetic network. The deformation vector *<sup>d</sup>*^ *C* thus obtained will not provide reliable data for testing the particular deformations in the subsidence.

The proposed and presented theory of the specific solution of the deformation vector in the case of any structural changes in the geodetic network will be acceptable for its proving in an analytical way, if we compare the deformation vector structures *<sup>d</sup>C*^ and *<sup>d</sup>*^ *C* expressed according to Eqs. (5) and (12). Then the structure of the deformation vector *<sup>d</sup>*^ *C* is expressed according to Eq. (12), and the further equation will be valid:

$$\begin{aligned} d\hat{\overline{\mathbf{C}}} &= \left[ \overline{\mathbf{C}}^{\theta} + \overline{\mathbf{G}} \left( \mathbf{L}\_{(i)} - \overline{\mathbf{L}}^{\theta} \right) \right] - \left[ \mathbf{C}^{\theta} + \mathbf{G} \left( \mathbf{L}\_{(0)} - \mathbf{L}^{\bullet} \right) \right] \\ &= \overline{\mathbf{G}} \left( \mathbf{L}\_{(i)} - \mathbf{L}^{\bullet} \right) - \mathbf{G} \left( \mathbf{L}\_{(0)} - \mathbf{L}^{\bullet} \right) + \overline{\mathbf{C}}^{\theta} - \mathbf{C}^{\theta} \end{aligned} \tag{13}$$

breakpoints determine the edges of the subsurface at which the naturally consistent

As already mentioned in the Introduction, the geometric and thus data structure of the geodetic network of the monitoring station in the subsidence may be changed by some external intervention, such as some unforeseen earthworks and construction works at the monitoring station. Estimation of the structures of geodetic networks based on the Gauss-Markov model is the most used and the most effective method for their processing. In determining the statistical formulation of the Gauss-Markov model, we start from the following equations [11, 12, 17, 30–33]:

*<sup>Σ</sup><sup>L</sup>* <sup>¼</sup> *<sup>σ</sup>*<sup>2</sup>

of the approximate 3D coordinate values, *L*ð Þ <sup>0</sup> is the vector of the approximate observation magnitude values of the observed elements in the first measuring epoch *<sup>t</sup>*ð Þ <sup>0</sup> , *<sup>L</sup><sup>o</sup>* is the vector of the approximate observation magnitude values of the observed elements, *dC*^ is the deformation vector, *dL* is the vector of the measured

values supplements, *Σ<sup>L</sup>* is the covariance matrix of the measured values, *σ*<sup>2</sup>

It will also be appeared in the changed structures, let us say in a size of the

presupposed model of a network adjustment following out from the Gauss-Markov

If between monitoring epochs there are no changes in the geometrical and observational structure of the geodetic network, then the matrixes and vectors

vector *dC*^ can be determined by a conventional procedure according to the

*ATQ*�**<sup>1</sup>**

3D coordinates of the observed points according to the equation:

• In the basic (first) monitoring epoch *<sup>t</sup>*ð Þ <sup>0</sup> , we have the vector *<sup>C</sup>*^ð Þ <sup>0</sup> of the

adjusted 3D coordinates of the observed points which are obtained according

• In the other following epochs *<sup>t</sup>*ð Þ*<sup>i</sup>* , we also obtain the vector *<sup>C</sup>*^ð Þ*<sup>i</sup>* of the adjusted

and *L<sup>o</sup>* remain identical for each epoch. Only in such case the deformation

priori variance, and *QL* is the cofactor matrix of the observations.

matrixes and vectors *A*, *QL*, *C<sup>o</sup>* and *Lo*

where *v* is the vector of corrections of the measured (observed) values *L*,*A* is the configuration (modeling) matrix of the geodetic network (otherwise called also the Jacobian matrix), i.e., the matrix of the partial derivatives of functions *Lo* <sup>¼</sup> *<sup>f</sup> <sup>C</sup><sup>o</sup>* ð Þ

,*C*^ is the vector of the aligned 3D coordinate values, *C<sup>o</sup>* is the vector

� *<sup>L</sup>*ð Þ **<sup>0</sup>** � *<sup>L</sup><sup>o</sup>* <sup>¼</sup> *<sup>A</sup>dC*^ � *<sup>d</sup><sup>L</sup>* (1)

. These matrixes and vectors enter into the

*<sup>L</sup> <sup>L</sup>*ð Þ **<sup>0</sup>** � *<sup>L</sup><sup>o</sup>* <sup>¼</sup> *<sup>C</sup><sup>o</sup>* <sup>þ</sup> *G L*ð Þ **<sup>0</sup>** � *<sup>L</sup><sup>o</sup>* (3)

<sup>0</sup>*Q<sup>L</sup>* (2)

<sup>0</sup> is a

coherent of the earth surface is broken and the subsidence begins to form. 3D deformation vector models help to support the location of the breakpoints

**2. Theory of the deformation vector specific solution**

*<sup>v</sup>* <sup>¼</sup> *<sup>A</sup> <sup>C</sup>*^ � *<sup>C</sup><sup>o</sup>*

[10–13, 18, 21].

*Mining Techniques - Past, Present and Future*

by the vector *C<sup>o</sup>*

model [11, 12, 29–31].

*A*, *QL*, *C<sup>o</sup>*

**138**

**2.1 Deformation vector**

following model [2, 16, 17]:

to the Gauss-Markov model:

*<sup>L</sup> <sup>A</sup>* �**<sup>1</sup>**

*<sup>C</sup>*^ð Þ **<sup>0</sup>** <sup>¼</sup> *<sup>C</sup><sup>o</sup>* <sup>þ</sup> *<sup>A</sup>TQ*�**<sup>1</sup>**

and on the base of Eqs. (6)–(9) and also on the base of the linearization of *G* into *<sup>G</sup>* <sup>¼</sup> *<sup>G</sup>* <sup>þ</sup> *<sup>d</sup>G*, the following derivation will be applied for the deformation vector *<sup>d</sup>*^ *C*:

$$\begin{aligned} \dot{d\overline{\mathbf{C}}} &= (\mathbf{G} + d\mathbf{G}) \left( \mathbf{L}\_{(i)} - \overline{\mathbf{L}}^{\sigma} \right) - \mathbf{G} \left( \mathbf{L}\_{(0)} - \mathbf{L}^{\sigma} \right) + d\mathbf{C}^{\sigma} = \overline{\mathbf{G}} \left[ \mathbf{L}\_{(i)} - \left( \mathbf{L}^{\sigma} + d\mathbf{L}^{\sigma} \right) \right] + d\mathbf{G} \left( \mathbf{L}\_{(i)} - \overline{\mathbf{L}}^{\sigma} \right) - \mathbf{G} \left( \mathbf{L}\_{(i)} - \mathbf{L}^{\sigma} \right) \\ - \mathbf{G} \left( \mathbf{L}\_{(0)} - \mathbf{L}^{\sigma} \right) + d\mathbf{C}^{\sigma} = \mathbf{G} \left( \mathbf{L}\_{(i)} - \mathbf{L}^{\sigma} \right) + \mathbf{G} d\mathbf{L}^{\sigma} + d\mathbf{G} \left( \mathbf{L}\_{(i)} - \mathbf{L}^{\sigma} \right) - \mathbf{G} \left( \mathbf{L}\_{(0)} - \mathbf{L}^{\sigma} \right) + d\mathbf{C}^{\sigma} = \\ \mathbf{G} \left( \mathbf{L}\_{(i)} - \mathbf{L}\_{(0)} \right) + \mathbf{G} d\mathbf{L}^{\sigma} + d\mathbf{G} \left( \mathbf{L}\_{(i)} - \mathbf{L}^{\sigma} \right) + d\mathbf{C}^{\sigma} \end{aligned} \tag{14}$$

and finally the deformation vector *<sup>d</sup>*^ *C* will be calculated according to the following equation:

$$d\hat{\mathbf{C}} = d\hat{\mathbf{C}} + \delta d\hat{\mathbf{C}} \tag{15}$$

because using the identical *C<sup>o</sup>* and *Lo* are not the problem to adhere in the individual epochs. Or the deformation vector corrections *δdC*^ are calculated according to Eqs. (10), (11), and (13), so that the deformation vector *dC*^ is then

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming…*

The monitoring station is situated in the territory of the mining field of the abandoned magnesite mine of Košice-Bankov. This territory was characterized by a devastated mining surface with many mining tailing piles and especially the large subsidence. The city district of Košice-Bankov is located in the northern part of the city of Košice. In addition to the abandoned magnesite mine, there is a very popular urban recreational and touristic resort, located in the large urban forest park of the city of Košice. The territory of the urban recreational and touristic resort and forest park are situated in close proximity, respectively, in the territory above the mining

*Ortho-photo map of the city of Košice with a detail view to the mine field of Košice-Bankov.*

corrected according to the introduced Eq. (15).

*DOI: http://dx.doi.org/10.5772/intechopen.91461*

field of the former magnesite mine (**Figure 1**).

**3. Study case example**

**Figure 1.**

**141**

**3.1 Study region description**

Eq. (15) notates that the deformation vector *<sup>d</sup>*^ *C* (calculated at some changes in the geodetic network structure) is different from its vector of the correct values *dC*^ only by the component *δdC*^ (i.e., the correction component of the deformation vector corrections). In such a set case, the component *δdC*^ is generated not only by the spatial movement of points in the geodetic network between the particular epochs of the geodetic measurements, but at the same time it is generated by changes in the geometric and data structure of the network between the particular epochs due to some changes in its point field.

To avoid the so-called degradation of the deformation vector *dC*^ due to changes in the geometric and data structure of the geodetic network and at the same time for the deformation vector to express the real spatial changes in the subsidence, the presented theory offers the following procedures:


$$\begin{aligned} d\hat{\mathbf{C}} &= \mathbf{C}^{\boldsymbol{o}} + \left(\mathbf{A}^{T}\mathbf{Q}\_{L}^{-1}\mathbf{A}\right)^{-1}\_{(i)}\mathbf{A}\_{(i)}^{T}\mathbf{Q}\_{L(i)}^{-1}\left(\mathbf{L}\_{(i)} - \mathbf{L}^{\boldsymbol{o}}\right) \\ &- \left[\mathbf{C}^{\boldsymbol{o}} + \left(\mathbf{A}^{T}\mathbf{Q}\_{L}^{-1}\mathbf{A}\right)^{-1}\_{(0)}\mathbf{A}\_{(0)}^{T}\mathbf{Q}\_{L(0)}^{-1}\left(\mathbf{L}\_{(0)} - \mathbf{L}^{\boldsymbol{o}}\right)\right] \end{aligned} \tag{16}$$

and

$$d\hat{\mathbf{C}} = \mathbf{G}\_{(i)}\mathbf{L}\_{(i)} - \mathbf{G}\_{(0)}\mathbf{L}\_{(0)} - \mathbf{L}^{\circ} \left(\mathbf{G}\_{(i)} - \mathbf{G}\_{(0)}\right) \tag{17}$$

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming… DOI: http://dx.doi.org/10.5772/intechopen.91461*

because using the identical *C<sup>o</sup>* and *Lo* are not the problem to adhere in the individual epochs. Or the deformation vector corrections *δdC*^ are calculated according to Eqs. (10), (11), and (13), so that the deformation vector *dC*^ is then corrected according to the introduced Eq. (15).

#### **3. Study case example**

and on the base of Eqs. (6)–(9) and also on the base of the linearization of *G* into *<sup>G</sup>* <sup>¼</sup> *<sup>G</sup>* <sup>þ</sup> *<sup>d</sup>G*, the following derivation will be applied for the deformation vector *<sup>d</sup>*^

*<sup>C</sup>* <sup>¼</sup> ð Þ *<sup>G</sup>* <sup>þ</sup> *<sup>d</sup><sup>G</sup> <sup>L</sup>*ð Þ*<sup>i</sup>* � *<sup>L</sup>* � �*<sup>o</sup>* � *G L*ð Þ <sup>0</sup> � *<sup>L</sup><sup>o</sup>* � � <sup>þ</sup> *<sup>d</sup>C<sup>o</sup>* <sup>¼</sup> *G L*ð Þ*<sup>i</sup>* � *<sup>L</sup><sup>o</sup>* <sup>þ</sup> *<sup>d</sup>L<sup>o</sup>* ð Þ � � <sup>þ</sup> *<sup>d</sup>G L*ð Þ*<sup>i</sup>* � *<sup>L</sup>* � �*<sup>o</sup>* � �*G L*ð Þ <sup>0</sup> � *<sup>L</sup><sup>o</sup>* � � <sup>þ</sup> *<sup>d</sup>C<sup>o</sup>* <sup>¼</sup> *G L*ð Þ*<sup>i</sup>* � *<sup>L</sup><sup>o</sup>* � � <sup>þ</sup> *<sup>G</sup>dL<sup>o</sup>* <sup>þ</sup> *<sup>d</sup>G L*ð Þ*<sup>i</sup>* � *<sup>L</sup><sup>o</sup>* � � � *G L*ð Þ <sup>0</sup> � *<sup>L</sup><sup>o</sup>* � � <sup>þ</sup> *<sup>d</sup>C<sup>o</sup>* <sup>¼</sup>

the geodetic network structure) is different from its vector of the correct values *dC*^ only by the component *δdC*^ (i.e., the correction component of the deformation vector corrections). In such a set case, the component *δdC*^ is generated not only by the spatial movement of points in the geodetic network between the particular epochs of the geodetic measurements, but at the same time it is generated by changes in the geometric and data structure of the network between the particular

To avoid the so-called degradation of the deformation vector *dC*^ due to changes in the geometric and data structure of the geodetic network and at the same time for the deformation vector to express the real spatial changes in the subsidence, the

• The geodetic networks at the monitoring stations shall be designed in order to achieve the maximal physical integrity of its points (object and especially reference points) throughout the entire monitoring period. When designing a monitoring station, expert consultation with representatives of a spatial

• If some reference points were lost or destroyed, new points should be stabilized in enough proximity to these lost or destroyed reference points as possible. The

• However, if the matrixes *A* and *QL* are significantly or even slightly changed between the monitoring epochs *t*ð Þ <sup>0</sup> and *t*ð Þ*<sup>i</sup>* (e.g., in *t*ð Þ <sup>0</sup> the geodetic network was measured by a trilateration measurement way, and in *t*ð Þ*<sup>i</sup>* by traverse measurement way, it is necessary to observe (measure) other new magnitudes, etc.), then the deformation vector *dC*^ can be determined according to the

> ð Þ*<sup>i</sup> <sup>A</sup><sup>T</sup>* ð Þ*<sup>i</sup> <sup>Q</sup>*�<sup>1</sup>

*<sup>d</sup>C*^ <sup>¼</sup> *<sup>G</sup>*ð Þ*<sup>i</sup> <sup>L</sup>*ð Þ*<sup>i</sup>* � *<sup>G</sup>*ð Þ <sup>0</sup> *<sup>L</sup>*ð Þ <sup>0</sup> � *<sup>L</sup><sup>o</sup> <sup>G</sup>*ð Þ*<sup>i</sup>* � *<sup>G</sup>*ð Þ <sup>0</sup>

ð Þ <sup>0</sup> *AT*

*<sup>L</sup> <sup>A</sup>* � ��<sup>1</sup>

*L i*ð Þ *<sup>L</sup>*ð Þ*<sup>i</sup>* � *<sup>L</sup><sup>o</sup>* � �

� � (17)

ð Þ **<sup>0</sup>** *<sup>Q</sup>*�**<sup>1</sup>** *<sup>L</sup>*ð Þ **<sup>0</sup>** *<sup>L</sup>*ð Þ <sup>0</sup> � *<sup>L</sup><sup>o</sup>* � � h i

planning and also with the mine district owners is essential.

*<sup>L</sup> <sup>A</sup>* � ��<sup>1</sup>

� *<sup>C</sup><sup>o</sup>* <sup>þ</sup> *ATQ*�**<sup>1</sup>**

*d*^

*C* will be calculated according to the

*<sup>C</sup>* <sup>¼</sup> *<sup>d</sup>C*^ <sup>þ</sup> *<sup>δ</sup>dC*^ (15)

*C* (calculated at some changes in

*d*^

¼ *G L*ð Þ*<sup>i</sup>* � *L*ð Þ <sup>0</sup>

following equation:

� � <sup>þ</sup> *<sup>G</sup>dL<sup>o</sup>* <sup>þ</sup> *<sup>d</sup>G L*ð Þ*<sup>i</sup>* � *<sup>L</sup>* � �*<sup>o</sup>* <sup>þ</sup> *<sup>d</sup>C<sup>o</sup>*

Eq. (15) notates that the deformation vector *<sup>d</sup>*^

and finally the deformation vector *<sup>d</sup>*^

*Mining Techniques - Past, Present and Future*

epochs due to some changes in its point field.

presented theory offers the following procedures:

same principle is held for the object points.

*<sup>d</sup>C*^ <sup>¼</sup> *Co* <sup>þ</sup> *<sup>A</sup><sup>T</sup>Q*�<sup>1</sup>

following equations:

and

**140**

*C*:

(14)

(16)

#### **3.1 Study region description**

The monitoring station is situated in the territory of the mining field of the abandoned magnesite mine of Košice-Bankov. This territory was characterized by a devastated mining surface with many mining tailing piles and especially the large subsidence. The city district of Košice-Bankov is located in the northern part of the city of Košice. In addition to the abandoned magnesite mine, there is a very popular urban recreational and touristic resort, located in the large urban forest park of the city of Košice. The territory of the urban recreational and touristic resort and forest park are situated in close proximity, respectively, in the territory above the mining field of the former magnesite mine (**Figure 1**).

**Figure 1.** *Ortho-photo map of the city of Košice with a detail view to the mine field of Košice-Bankov.*

Until the 1970s of the twentieth century, any systematic attention was not paid to the extent of mining damage on the earth surface at the territory of Košice-Bankov as a result of magnesite mining. It was only after this period that scientific studies began to be taken into account when dealing with the creation of the large subsidence and devastation of the protected area of the forest park and the environmental protection in the tangent territory of Košice-Bankov. The gradual development of the subsidence in the mining area of the magnesite mine of Košice-Bankov in the east of Slovakia has been monitored since the end of the 1970s by systematic geodetic measurements. The monitoring station project to monitor the development of the subsidence in the territory of Košice-Bankov was designed and implemented by researchers of the Technical University of Košice in 1976, when the first geodetic measurements were carried out and later by researchers of the Pavol Jozef Šafárik University in Košice. The first observed data were obtained from this monitoring station in autumn 1976, and since then the regular periodic spring and autumn geodetic terrestrial and satellite (GPS, GNSS) measurements were performed every year.

Before reclaiming the mining landscape on the territory of Košice-Bankov, the monitoring station was located on the site of the former subsidence at the mining shaft, which was called the Western shaft. The monitoring station was built from the geodetic network consisting of the network of the reference points (No: 01A, 01B, 01C, 01D) and the network of the object points (78 points in total). The object points were geometrically grouped into six geodetic network profiles (0–V) (**Figure 2**). All geodetic network profiles of the monitoring station of Košice-Bankov were geometrically spaced across and along the expected movements in the subsidence (**Figure 2**). Gradually, by creating the subsidence, some object points were destroyed by the nature destructive processes in the subsidence. **Figures 3** and **4** show the panoramic views to the subsidence of Košice-Bankov from the southwestern edge of this subsidence in 2001 and 2002. In that time the magnesite mine had been out of its operation for 3 years.

3D data (elements) of geodetic network points of the monitoring station were initially (since 1076) measured by 3D geodetic measurements (position and leveling measurements) by the application of the classical terrestrial geodetic technologies using the optical geodetic theodolites, electro-optical total stations, and leveling devices for a very precise leveling. Later, since 1977, the periodic measurements at the monitoring station have been made by the satellite geodetic methods GPS and GNSS, i.e., Trimble 3303DR Total Station, GPS: ProMark2, and GNSS: Leica Viva GS08.

In both geodetic periodic measurement technologies (terrestrial and satellite), regular geodetic measurements were performed twice a year, i.e., during the spring and autumn months [12, 17]. In 1981, some of the object points of the geodetic network of the monitoring station were damaged, respectively; the points were completely destroyed by some unplanned and uncontrolled earthworks in the vicinity of the subsidence (points No. 2, 3, 30, 38, 104, and 105 and 227<sup>1</sup> on the profiles No. 0, I, and II). Most of the damage or destruction of the abovementioned points occurred during the adjustment of some forest stands in the nearby forest park and by some earthworks on the surrounding mining tailing piles.

#### **3.2 Accuracy and quality assessment of the geodetic network**

1D, 2D, and 3D accuracy of the geodetic network points of the monitoring station was evaluated by testing the global and local network indices. The global indices were numerically expressed to assess the accuracy of the entire geodetic network.

For the global indices, the following were tested: *tr* (*ΣC*^ ), i.e., the track of the covariance matrix *ΣC*^ , the volume global indices, and det(*ΣC*^ ), i.e., the determinant

*The monitoring station of Košice-Bankov (reference points 01C and 01D destroyed points).*

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming…*

*DOI: http://dx.doi.org/10.5772/intechopen.91461*

In fact, the local indices were the point indices that characterize the reliability of

ffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffi

, the mean

*σ*2 *X*^*i* <sup>þ</sup> *<sup>σ</sup>*<sup>2</sup> *Y*^*i* <sup>þ</sup> *<sup>σ</sup>*<sup>2</sup> *Z*^*i*

q

of the covariance matrix.

**Figure 2.**

**Figure 3.**

**143**

the geodetic network points: the mean 3D error *σ<sup>p</sup>* ¼

*The subsidence of Košice-Bankov; panoramic view: Autumn 2001.*

<sup>1</sup> The reference point No. 227 (profile II) was rebuilt instead of the point No. 226.

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming… DOI: http://dx.doi.org/10.5772/intechopen.91461*

#### **Figure 2.**

Until the 1970s of the twentieth century, any systematic attention was not paid to the extent of mining damage on the earth surface at the territory of Košice-Bankov as a result of magnesite mining. It was only after this period that scientific studies began to be taken into account when dealing with the creation of the large subsidence and devastation of the protected area of the forest park and the environmental protection in the tangent territory of Košice-Bankov. The gradual development of the subsidence in the mining area of the magnesite mine of Košice-Bankov in the east of Slovakia has been monitored since the end of the 1970s by systematic geodetic measurements. The monitoring station project to monitor the development of the subsidence in the territory of Košice-Bankov was designed and implemented by researchers of the Technical University of Košice in 1976, when the first geodetic measurements were carried out and later by researchers of the Pavol Jozef Šafárik University in Košice. The first observed data were obtained from this monitoring station in autumn 1976, and since then the regular periodic spring and autumn geodetic terrestrial and satellite

Before reclaiming the mining landscape on the territory of Košice-Bankov, the monitoring station was located on the site of the former subsidence at the mining shaft, which was called the Western shaft. The monitoring station was built from the geodetic network consisting of the network of the reference points (No: 01A, 01B, 01C, 01D) and the network of the object points (78 points in total). The object

points were geometrically grouped into six geodetic network profiles (0–V) (**Figure 2**). All geodetic network profiles of the monitoring station of Košice-Bankov were geometrically spaced across and along the expected movements in the subsidence (**Figure 2**). Gradually, by creating the subsidence, some object points were destroyed by the nature destructive processes in the subsidence. **Figures 3** and **4** show the panoramic views to the subsidence of Košice-Bankov from the southwestern edge of this subsidence in 2001 and 2002. In that time the magnesite mine had

3D data (elements) of geodetic network points of the monitoring station were initially (since 1076) measured by 3D geodetic measurements (position and leveling measurements) by the application of the classical terrestrial geodetic technologies using the optical geodetic theodolites, electro-optical total stations, and leveling devices for a very precise leveling. Later, since 1977, the periodic measurements at the monitoring station have been made by the satellite geodetic methods GPS and GNSS, i.e., Trimble 3303DR Total Station, GPS: ProMark2, and GNSS: Leica Viva GS08. In both geodetic periodic measurement technologies (terrestrial and satellite), regular geodetic measurements were performed twice a year, i.e., during the spring and autumn months [12, 17]. In 1981, some of the object points of the geodetic network of the monitoring station were damaged, respectively; the points were completely destroyed by some unplanned and uncontrolled earthworks in the vicinity of the subsidence (points No. 2, 3, 30, 38, 104, and 105 and 227<sup>1</sup> on the profiles No. 0, I, and II). Most of the damage or destruction of the abovementioned points occurred during the adjustment of some forest stands in the nearby forest

park and by some earthworks on the surrounding mining tailing piles.

numerically expressed to assess the accuracy of the entire geodetic network.

1D, 2D, and 3D accuracy of the geodetic network points of the monitoring station was evaluated by testing the global and local network indices. The global indices were

**3.2 Accuracy and quality assessment of the geodetic network**

<sup>1</sup> The reference point No. 227 (profile II) was rebuilt instead of the point No. 226.

(GPS, GNSS) measurements were performed every year.

*Mining Techniques - Past, Present and Future*

been out of its operation for 3 years.

**142**

*The monitoring station of Košice-Bankov (reference points 01C and 01D destroyed points).*

#### **Figure 3.** *The subsidence of Košice-Bankov; panoramic view: Autumn 2001.*

For the global indices, the following were tested: *tr* (*ΣC*^ ), i.e., the track of the covariance matrix *ΣC*^ , the volume global indices, and det(*ΣC*^ ), i.e., the determinant of the covariance matrix.

In fact, the local indices were the point indices that characterize the reliability of the geodetic network points: the mean 3D error *σ<sup>p</sup>* ¼ ffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffi *σ*2 *X*^*i* <sup>þ</sup> *<sup>σ</sup>*<sup>2</sup> *Y*^*i* <sup>þ</sup> *<sup>σ</sup>*<sup>2</sup> *Z*^*i* q , the mean

**Figure 4.** *The subsidence of Košice-Bankov; panoramic view: Spring 2002.*

coordinate error *σXYZ* ¼ ffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffi *σ*2 *X*^*i* <sup>þ</sup>*σ*<sup>2</sup> *Y*^*i* <sup>þ</sup>*σ*<sup>2</sup> *Z*^ *i* 3 r , and the confidence absolute ellipses or ellipsoids, which were used to assess the real 2D or 3D point accuracy. It is necessary to know the design elements of the ellipse of errors, i.e., the semi-major axis *a*, the semi-minor axis *b*, the bearing *φ<sup>a</sup>* of the semi-major axis, and the ellipsoid flattening *f* (*f* ¼ 1 � *b=a*).

**Point** *ai* **(mm)** *bi* **(mm)** *φai* **(gon)** *f* 49.9/51.0 5.9/5.1 172.303/172.695 1.8818/0.9 40.8/32.5 12.3/3.7 172.704/179.151 0.6985/0.8862 43.0/45.9 18.2/21.5 160.340/160.058 0.5767/0.5316 23.5/28.1 21.8/22.4 40.966/41.203 0.0723/0.2028 47.5/79.2 24.0/9.9 211.146/217.148 0.4947/0.875 42.8/42.4 15.3/17.6 370.337/370.624 0.6425/0.5849 28.8/25.2 8.1/10.9 19.634/19.781 0.7188/0.5675

> **Determinant** *det(ΣC*^ *)*

2.869.1025/ 2.869.10<sup>25</sup>

 36.4/37.5 25.7/18.8 30.9/28.4 21.8/24.5 33.9/33.0 23.9/23.0 23.3/26.7 16.5/13.8 38.6/14.1 27.3/54.9 32.9/27.4 23.3/19.9 21.7/22.5 15.3/17.4

**Point Mean 3D error** *σ<sup>p</sup>* **(mm) Mean coordinate error** *σXYZ* **(mm)**

**Average mean error** *σC*^pr **[mm]**

22.428/22.971 124.218/

**Norm** *nor dC*^ 

**[mm]**

126.155

**Point** *mX* **(mm)** *mY* **(mm)** *mZ* **(mm)** 15.7/16.4 32.9/45.5 12.5/72.4 14.8/34.3 27.2/58.9 30.5/69.1 21.1/25.6 26.5/24.1 45.5/32.7 16.6/14.9 16.3/8.1 20.1/18.4 18.2/41.3 34.1/69.0 55.4/78.0 28.2/31.6 17.1/21.1 9.9/17.8 20.0/16.9 8.5/4.7 10.9/10.8

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming…*

**Table 2.**

**Table 3.**

**Table 4.**

**145**

*Local indices (1976/2014).*

**Table 1.**

*Mean errors (1976/2014).*

*DOI: http://dx.doi.org/10.5772/intechopen.91461*

*Absolute confidence ellipse elements (*α *= 0.05) (1976/2014).*

**[mm2 ]**

7041.054

**Rank** rk*(ΣC*^ *)* **Track** tr*(ΣC*^ *)*

14/14 7041.901/

*Global indices (1976/2014).*

The geodetic network quality characteristics to be assessed are, above all, the accuracy and reliability of the position of the points. In addition to numerical expressions, the accuracy of the point position can also be expressed using the graphical indicators such as the reliability curves and the ellipse of confidence (ellipsoids of reliability in the case of 3D space). The ellipsoids determine the random space in which the actual location of the points will lie with a probability of *1-α*, where *α* is the level of significance chosen according to which the ellipsoids are of different size. For 3D space in a geodetic practice, the standard confidence ellipsoids are usually used. The design parameters of such ellipsoids can be derived either from the cofactor matrix *QL*of the adjusted coordinates, where the design parameters are arranged on the main diagonal of the matrix, or from the covariance matrix of the coordinate estimations *ΣC*^ of the determined points, which are arranged also on the main diagonal of that matrix.

All calculated data according to the submitted specific theory of the deformation vector solution in the case of any geometric and data changes in the geodetic network of the monitoring station are shown in **Tables 1**–**5** which are focused on the accuracy and quality assessment of the network in the years 1976 and 2014 (1976/2014<sup>2</sup> ) (**Tables 1**–**5**). **Tables 1**–**5** comprehends the adjusted mean errors of the individual coordinates, global and local 3D indices, and their absolute confidence ellipsoid elements determining 3D accuracy of some chosen replaced points; the numbers in front of the slash belong to 1976; the numbers after the forward slash belong to 2014. In 2007 the points No. 2, 3, 30, 38, 104, 105, and 227 were restabilized due to small earthworks for the purpose of some preparation work for the future reclaiming of the mining territory of Košice-Bankov. The values of the deformation vectors confirm the fact that the presented theory of the deformation vector specific solution is suitable and variable adapted to various damages of geodetic networks [7]. In geodetic practice, however, there are cases where the values of deformation vectors need not mean any displacement of the geodetic network points (movement around the point of the geodetic network). Although the geodetic network points were adjusted according to the conventional method using the Gauss-Mark model, the deformation vector values may be loaded by the accumulation of some measurement errors. For this reason, when evaluating the

<sup>2</sup> Deformation survey on the monitoring station of Košice-Bankov without the reclaiming work intervention was finished in the autumn 2014.

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming… DOI: http://dx.doi.org/10.5772/intechopen.91461*


#### **Table 1.**

coordinate error *σXYZ* ¼

*f* (*f* ¼ 1 � *b=a*).

**Figure 4.**

(1976/2014<sup>2</sup>

**144**

ffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffi *σ*2 *X*^*i* <sup>þ</sup>*σ*<sup>2</sup> *Y*^*i* <sup>þ</sup>*σ*<sup>2</sup> *Z*^ *i* 3

soids, which were used to assess the real 2D or 3D point accuracy. It is necessary to know the design elements of the ellipse of errors, i.e., the semi-major axis *a*, the semi-minor axis *b*, the bearing *φ<sup>a</sup>* of the semi-major axis, and the ellipsoid flattening

The geodetic network quality characteristics to be assessed are, above all, the accuracy and reliability of the position of the points. In addition to numerical expressions, the accuracy of the point position can also be expressed using the graphical indicators such as the reliability curves and the ellipse of confidence (ellipsoids of reliability in the case of 3D space). The ellipsoids determine the random space in which the actual location of the points will lie with a probability of *1-α*, where *α* is the level of significance chosen according to which the ellipsoids are of different size. For 3D space in a geodetic practice, the standard confidence ellipsoids are usually used. The design parameters of such ellipsoids can be derived either from the cofactor matrix *QL*of the adjusted coordinates, where the design parameters are arranged on the main diagonal of the matrix, or from the covariance

matrix of the coordinate estimations *ΣC*^ of the determined points, which are

vector solution in the case of any geometric and data changes in the geodetic network of the monitoring station are shown in **Tables 1**–**5** which are focused on the accuracy and quality assessment of the network in the years 1976 and 2014

the individual coordinates, global and local 3D indices, and their absolute confidence ellipsoid elements determining 3D accuracy of some chosen replaced points; the numbers in front of the slash belong to 1976; the numbers after the forward slash belong to 2014. In 2007 the points No. 2, 3, 30, 38, 104, 105, and 227 were restabilized due to small earthworks for the purpose of some preparation work for the future reclaiming of the mining territory of Košice-Bankov. The values of the deformation vectors confirm the fact that the presented theory of the deformation vector specific solution is suitable and variable adapted to various damages of geodetic networks [7]. In geodetic practice, however, there are cases where the values of deformation vectors need not mean any displacement of the geodetic network points (movement around the point of the geodetic network). Although the geodetic network points were adjusted according to the conventional method using the Gauss-Mark model, the deformation vector values may be loaded by the accumulation of some measurement errors. For this reason, when evaluating the

<sup>2</sup> Deformation survey on the monitoring station of Košice-Bankov without the reclaiming work

All calculated data according to the submitted specific theory of the deformation

) (**Tables 1**–**5**). **Tables 1**–**5** comprehends the adjusted mean errors of

, and the confidence absolute ellipses or ellip-

r

*The subsidence of Košice-Bankov; panoramic view: Spring 2002.*

*Mining Techniques - Past, Present and Future*

arranged also on the main diagonal of that matrix.

intervention was finished in the autumn 2014.

*Mean errors (1976/2014).*


#### **Table 2.**

*Absolute confidence ellipse elements (*α *= 0.05) (1976/2014).*


#### **Table 3.**

*Global indices (1976/2014).*


**Table 4.** *Local indices (1976/2014).*


**Table 5.**

*Deformation vector values (1976/2014).*

significance of deformation vectors, it is necessary to test them by means of the global and localization test of a congruence (see Chapter 3.3). After a series of the last geodetic measurements in spring 2014, the deformation vectors at the tested points No. 2, 3, 30, 38, 104, 105, and 227 of the geodetic network of the monitoring station have moved from �4 to 9.7 mm (**Table 5**). 3D mean errors ranged from 14.1 to 38.6 mm, and the mean coordinate errors were from 13.8 to 54.9 mm (**Table 4**). In autumn 2014 all points of the monitoring station of Košice-Bankov were destroyed by the reclaiming work, i.e., the reference points were removed and the object points were backfilled by a secondary imported soil.

#### **3.3 Global test of the geodetic network congruence**

Significant stability, respectively instability of the geodetic network points, is rejected or not rejected by verifying the null-hypothesis *H*<sup>0</sup> and also other alternative hypothesis [11, 12, 17, 34].

$$H\_0: d\hat{\overline{\mathbf{C}}} = 0; \quad H\_a: d\hat{\overline{\mathbf{C}}} \neq 0 \tag{18}$$

case, it can be stated that the deformation occurred with the level *α* of a

*Test statistics results of the geodetic network points at the monitoring station of Košice-Bankov.*

**Point** *TG i*ð Þ **< ≤ >** *F* **Notice**

2 1.297 < 3.724 Deformation vectors are not significant

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming…*

**Table 6** presents the global tested results of the geodetic network congruence.

The Geographical Information Systems (GIS) of the mining landscape of Košice-Bankov is based on the following key points [25]: basic and simple presentation of geodata, management of the basic database, and wide availability of information. The best feasible solution for the implementation of the GIS project is the Free Open Source software applications, which are easily available on the Internet. The general function of the Free Open Source software application is the viability of free code and data sources via HTTP and FTP protocols located on the project website. Other features of the Free Open Source range include ease of use, access to data and information, centralized system configuration, modular things, and any Open Source platform (depending on PHP, MySQL, and ArcIMS ports) [7, 25, 35–39]. The network application MySQL is currently the most advantageous database system on the Internet and was also applied to the deformation survey outputs from

Whole database part in GIS for the subsidence of Košice-Bankov in all applications was processed into the MySQL database (**Figure 5**). 3D model of the subsi-

implemented into the reclaiming plan of the mining territory of Košice-Bankov for

Given the fact that extraction of magnesite mineral has been completed at the mine of Košice-Bankov and this mine is abandoned since the end of the 90 years of the twentieth century and the investigation concluded that the Košice-Bankov with the huge subsidence is stable at the end of the deformation investigation, the municipality of the city of Košice (Department of land planning of the city) has adopted a definitive plan for reclaiming this mining landscape. Numerical and graphical analyses of the results from the long-term geodetic measurements of the deformations in the subsidence at the monitoring station of Košice-Bankov with their subsequent testing analyses of the deformation vectors confirmed the stability not only in the subsidence but also in the surrounding mining landscape. The subsidence and other mining earthworks of huge dimensions destroyed the entire surroundings of the mining plant (mining tailings piles, various excavations in the

dence of Košice-Bankov with the multilayered GIS applications has been

**4. Subsidence in GIS for reclaiming mining landscape**

the monitoring station of Košice-Bankov.

the needs of the municipality of the city of Košice.

reliability.

**Table 6.**

**147**

3 3.7236 ≤ 30 3.501 < 38 3.7237 ≤ 104 2.871 < 105 1.403 < 227 2.884 <

*DOI: http://dx.doi.org/10.5772/intechopen.91461*

where *H*<sup>0</sup> expresses insignificance of the coordinate differences (deformation vector) between epochs *t*ð Þ <sup>0</sup> and *t*ð Þ*<sup>i</sup>* .

The test statistics *TG* can be used for a global test:

$$T\_G = \frac{d\hat{\overline{\mathbf{C}}} \mathbf{Q}\_{d\hat{\overline{\mathbf{C}}}}^{-1} d\hat{\overline{\mathbf{C}}}^T}{k \overline{\mathbf{s}}\_0^2} \approx F(f\_1, f\_2) \tag{19}$$

where *<sup>Q</sup>d*^ *C* is the cofactor matrix of the final deformation vector *<sup>d</sup>*^ *C*, *k* is the coordinate number entering the geodetic network adjustment (*k*=3 for 3D coordinates), and *s* 2 <sup>0</sup> is the posteriori variation factor common for both epochs *t*ð Þ <sup>0</sup> and *t*ð Þ*<sup>i</sup>* .

The critical value *TKRIT* is searched in the tables of *F* distribution (the Fisher-Snedecor distribution) according to the degrees of freedom *f* <sup>1</sup> ¼ *f* <sup>2</sup> ¼ *n* � *k* or *f* <sup>1</sup> ¼ *f* <sup>2</sup> ¼ *n* � *k* þ *d*, where *n* is number of the measured values entering into the network adjustment and *d* is the network defect at the network free adjustment. Through the use of methods, MINQUE is *s* 2*t* ð Þ 0 <sup>0</sup> ¼ *s* 2*t* ð Þ*i* <sup>0</sup> ¼ *s* 2 <sup>0</sup> ¼ 1 [2, 11, 12, 16, 17, 34]. Test statistics *T* should be compared to critical test statistics *TKRIT*.*TKRIT* is found in the distribution Tables *F* according to the degrees of freedom of the geodetic network.

When comparing test statistics, there may be two cases:


*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming… DOI: http://dx.doi.org/10.5772/intechopen.91461*


#### **Table 6.**

significance of deformation vectors, it is necessary to test them by means of the global and localization test of a congruence (see Chapter 3.3). After a series of the last geodetic measurements in spring 2014, the deformation vectors at the tested points No. 2, 3, 30, 38, 104, 105, and 227 of the geodetic network of the monitoring station have moved from �4 to 9.7 mm (**Table 5**). 3D mean errors ranged from 14.1 to 38.6 mm, and the mean coordinate errors were from 13.8 to 54.9 mm (**Table 4**).

**C [mm] 2 3 30 38 104 105 227**

**Point**

2.4 �2.9 �8.0 6.7 �4.0 0.6 9.7

In autumn 2014 all points of the monitoring station of Košice-Bankov were destroyed by the reclaiming work, i.e., the reference points were removed and the

Significant stability, respectively instability of the geodetic network points, is rejected or not rejected by verifying the null-hypothesis *H*<sup>0</sup> and also other alterna-

*<sup>C</sup>* <sup>¼</sup> 0; *<sup>H</sup><sup>α</sup>* : *<sup>d</sup>*^

where *H*<sup>0</sup> expresses insignificance of the coordinate differences (deformation

is the cofactor matrix of the final deformation vector *<sup>d</sup>*^

coordinate number entering the geodetic network adjustment (*k*=3 for 3D coordi-

The critical value *TKRIT* is searched in the tables of *F* distribution (the Fisher-Snedecor distribution) according to the degrees of freedom *f* <sup>1</sup> ¼ *f* <sup>2</sup> ¼ *n* � *k* or *f* <sup>1</sup> ¼ *f* <sup>2</sup> ¼ *n* � *k* þ *d*, where *n* is number of the measured values entering into the network adjustment and *d* is the network defect at the network free adjustment. Through the

*T* should be compared to critical test statistics *TKRIT*.*TKRIT* is found in the distribu-

1.*TG* ≤ *TKRIT*: The null-hypothesis *H*<sup>0</sup> is accepted. That is, the differences in

2.*TG*>*TKRIT*: The null-hypothesis *H*<sup>0</sup> is refused. This means that the differences in coordinate values (deformation vectors) are statistically significant. In this

tion Tables *F* according to the degrees of freedom of the geodetic network.

coordinate values (i.e., deformation vectors) are not significant.

≈*F f* <sup>1</sup>, *f* <sup>2</sup>

<sup>0</sup> is the posteriori variation factor common for both epochs *t*ð Þ <sup>0</sup> and *t*ð Þ*<sup>i</sup>* .

*C* 6¼ 0 (18)

(19)

<sup>0</sup> ¼ 1 [2, 11, 12, 16, 17, 34]. Test statistics

*C*, *k* is the

object points were backfilled by a secondary imported soil.

*<sup>H</sup>*<sup>0</sup> : *<sup>d</sup>*^

The test statistics *TG* can be used for a global test:

*TG* ¼

2*t* ð Þ 0 <sup>0</sup> ¼ *s* 2*t* ð Þ*i* <sup>0</sup> ¼ *s* 2

When comparing test statistics, there may be two cases:

*d*^ *CQ*�<sup>1</sup> *d*^ *C d*^ *C T*

> *ks* 2 0

**3.3 Global test of the geodetic network congruence**

tive hypothesis [11, 12, 17, 34].

*Deformation vector values (1976/2014).*

*Mining Techniques - Past, Present and Future*

vector) between epochs *t*ð Þ <sup>0</sup> and *t*ð Þ*<sup>i</sup>* .

where *<sup>Q</sup>d*^

nates), and *s*

**146**

*d*^

**Table 5.**

*C*

2

use of methods, MINQUE is *s*

*Test statistics results of the geodetic network points at the monitoring station of Košice-Bankov.*

case, it can be stated that the deformation occurred with the level *α* of a reliability.

**Table 6** presents the global tested results of the geodetic network congruence.

#### **4. Subsidence in GIS for reclaiming mining landscape**

The Geographical Information Systems (GIS) of the mining landscape of Košice-Bankov is based on the following key points [25]: basic and simple presentation of geodata, management of the basic database, and wide availability of information. The best feasible solution for the implementation of the GIS project is the Free Open Source software applications, which are easily available on the Internet. The general function of the Free Open Source software application is the viability of free code and data sources via HTTP and FTP protocols located on the project website. Other features of the Free Open Source range include ease of use, access to data and information, centralized system configuration, modular things, and any Open Source platform (depending on PHP, MySQL, and ArcIMS ports) [7, 25, 35–39]. The network application MySQL is currently the most advantageous database system on the Internet and was also applied to the deformation survey outputs from the monitoring station of Košice-Bankov.

Whole database part in GIS for the subsidence of Košice-Bankov in all applications was processed into the MySQL database (**Figure 5**). 3D model of the subsidence of Košice-Bankov with the multilayered GIS applications has been implemented into the reclaiming plan of the mining territory of Košice-Bankov for the needs of the municipality of the city of Košice.

Given the fact that extraction of magnesite mineral has been completed at the mine of Košice-Bankov and this mine is abandoned since the end of the 90 years of the twentieth century and the investigation concluded that the Košice-Bankov with the huge subsidence is stable at the end of the deformation investigation, the municipality of the city of Košice (Department of land planning of the city) has adopted a definitive plan for reclaiming this mining landscape. Numerical and graphical analyses of the results from the long-term geodetic measurements of the deformations in the subsidence at the monitoring station of Košice-Bankov with their subsequent testing analyses of the deformation vectors confirmed the stability not only in the subsidence but also in the surrounding mining landscape. The subsidence and other mining earthworks of huge dimensions destroyed the entire surroundings of the mining plant (mining tailings piles, various excavations in the

#### **Figure 5.**

working of the earth surface, and other mining works in the surroundings of the former mine plant) were gradually filled with imported secondary soil. Based on the results of the extensive geodetic measurements of deformations of the mining subsidence and its surroundings in the destroyed mining area of Košice-Bankov, reclaiming works began at the beginning of this century. Some recent reclaiming works were completed in the summer of 2016.

In the territory of the former large subsidence, the new forest park was built as an environmental forest greenery in the part of the Košice-Bankov urban recreation and touristic area of the city of Košice. The subsidence was filled with imported natural and secondary materials from many construction works and earthworks in the city of Košice and its surroundings. Given that the subsidence was of huge proportions, such sporadic embankment works took too long, more than 10 years. After the embankment and other earthworks were completed, the new forest park was planted in the area of the former subsidence (**Figure 6**). Especially birch was planted. Birch trees are known for their rapid growth, and their root system is not demanding on the underlying soil. At present, the birch grove represents almost 5 years of a healthy forest park. The reconstruction of the recreational and touristic area of Košice-Bankov was completed in spring 2016 (**Figure 7**). The mining tailing piles and the entire ruined area of the former mining plant were also reclaimed.

Many solar collectors have been built on the places of the former mining tailing piles, which contribute to the renewable sources of electricity for the city of

*The reconstructed recreation and touristic zone and revitalized forest park after reclaiming the mining*

*The subsidence of Košice-Bankov after reclaiming; panoramic view—Summer 2016. Solar collectors on the places of the former mining tailing piles; new forest park in the background on the places of the former subsidence.*

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming…*

*DOI: http://dx.doi.org/10.5772/intechopen.91461*

Košice (**Figure 6**).

*landscape of Košice-Bankov.*

**Figure 7.**

**149**

**Figure 6.**

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming… DOI: http://dx.doi.org/10.5772/intechopen.91461*

#### **Figure 6.**

*The subsidence of Košice-Bankov after reclaiming; panoramic view—Summer 2016. Solar collectors on the places of the former mining tailing piles; new forest park in the background on the places of the former subsidence.*

#### **Figure 7.**

working of the earth surface, and other mining works in the surroundings of the former mine plant) were gradually filled with imported secondary soil. Based on the results of the extensive geodetic measurements of deformations of the mining subsidence and its surroundings in the destroyed mining area of Košice-Bankov, reclaiming works began at the beginning of this century. Some recent reclaiming

In the territory of the former large subsidence, the new forest park was built as an environmental forest greenery in the part of the Košice-Bankov urban recreation and touristic area of the city of Košice. The subsidence was filled with imported natural and secondary materials from many construction works and earthworks in the city of Košice and its surroundings. Given that the subsidence was of huge proportions, such sporadic embankment works took too long, more than 10 years. After the embankment and other earthworks were completed, the new forest park was planted in the area of the former subsidence (**Figure 6**). Especially birch was planted. Birch trees are known for their rapid growth, and their root system is not demanding on the underlying soil. At present, the birch grove represents almost 5 years of a healthy forest park. The reconstruction of the recreational and touristic area of Košice-Bankov was completed in spring 2016 (**Figure 7**). The mining tailing piles and the entire ruined area of the former mining plant were also reclaimed.

works were completed in the summer of 2016.

*ArcView user interface visualization of the subsidence.*

*Mining Techniques - Past, Present and Future*

**Figure 5.**

**148**

*The reconstructed recreation and touristic zone and revitalized forest park after reclaiming the mining landscape of Košice-Bankov.*

Many solar collectors have been built on the places of the former mining tailing piles, which contribute to the renewable sources of electricity for the city of Košice (**Figure 6**).

#### **5. Conclusions**

The determination of the deformation vectors from the differences between the adjusted geodetic point coordinate vectors obtained from at least two monitoring epochs is readily achievable if the original geometric and data structure of the geodetic network of the monitoring station between the individual monitoring epochs is strictly preserved. The presented research study provides both the theoretical and practical results of the possibility of solving deformation vectors in the geodetic network of the monitoring station in the case of its geometrical and data structure violation during the period of monitoring the movements of the earth surface, i.e., if the points of the geodetic network were damaged or completely destroyed between the different epochs of measurements, i.e., the geodetic network was nonhomogenous. The deformation vectors solved in accordance with the proposed specific theory of solving monitored deformations of the earth surface in the case of disturbance of the geodetic network homogeneity of the monitoring station provide via 3D models in GIS a reliable idea of spatial changes in the coordinates of the geodetic network points. The proposed theory of the specific deformation vector solution leads to a reliable support in investigation of various deformations of the earth surface, such as mining subsidence, landslides, geotectonic (recent movements), movements of dams, and other important building objects.

3D models of the mining subsidence in GIS are useful tools for many reclaiming works in restoring ecosystems with some essential elements of the security measures against the possible and unforeseen consequences of former mining activities for the protection of the health and life of people moving in various mining areas.

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming…*

This research study was supported by the Slovak Research and Development Agency under the contract No. SK-CN-RD-18-0015 and by the European Union within the Interreg ENI CBC Programme under the contract No. HUSKROUA/ 1702/8.1/0065 and by the Scientific Grant Agency of the Ministry of Education, Science, Research and Sports of the Slovak Republic under the contract No. VEGA

**Acknowledgements**

*DOI: http://dx.doi.org/10.5772/intechopen.91461*

1/0839/18.

**Author details**

Vladimír Sedlák

**151**

Pavol Jozef Šafárik University in Košice, Slovakia

provided the original work is properly cited.

\*Address all correspondence to: vladimir.sedlak@upjs.sk

© 2020 The Author(s). Licensee IntechOpen. This chapter is distributed under the terms of the Creative Commons Attribution License (http://creativecommons.org/licenses/ by/3.0), which permits unrestricted use, distribution, and reproduction in any medium,

The largest values of the deformation vectors given in **Table 5** occurred at points No. 30 and No. 227. However, due to the fact that the deformation vectors tested at these points were not significant according to the test statistics, we can declare these points as static. The mentioned study case from the mining territory of the abandoned magnesite mine of Košice-Bankov confirmed the availability and applicability of the presented specific theory in the solution of the deformation vector in the deformation monitoring in the mining subsidence, where several violations of the geometric and data structure (homogeneity) of the geodetic network of the monitoring station occurred. Despite the validated method for the specific solution of the deformation vector in the geometric and data inhomogeneity of the geodetic network at the monitoring station, it should be stated that it would be preferable to maintain the homogeneity of the geodetic network during the whole monitoring period. Maintaining the homogeneity of the data of the geodetic network structure can ensure the permanent stabilization of the network points and the correct technically and physically implemented protection of the whole monitoring station from unexpected external interventions into such a station.

3D model situations of the mining subsidence in GIS platform from the mining territory of Košice-Bankov were delivered to the municipality of the city of Košice (especially for the Department of Land Planning and Chief Architect of the City) to deal with a spatial planning for the future environmental reclaiming of this abandoned mining region, such as the magnesite mines of Košice-Bankov. The analysis of the deformation vectors at the geodetic network points of the monitoring station located in the mining subsidence and in the surrounding mining territory of the abandoned magnesite mine of Košice-Bankov was important in defining and specifying the subsidence edges and the subsidence zones with a number of dangerous cracks and fissures. The very precise identification of 3D position of such delimitation of the mining subsidence constituted the basic document for the plan of the municipality of the city of Košice for the reclaiming of the entire devastated mining landscape of Košice-Bankov. The revitalization of the Košice-Bankov recreational and tourist zone with the adjacent urban forest park has been achieved through the comprehensive reclaiming that devastated mining landscape. The variability of 3D models of the mining subsidence allowed a wide spectrum at modeling of natural and also industrial disasters in the former mining territory of Košice-Bankov.

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming… DOI: http://dx.doi.org/10.5772/intechopen.91461*

3D models of the mining subsidence in GIS are useful tools for many reclaiming works in restoring ecosystems with some essential elements of the security measures against the possible and unforeseen consequences of former mining activities for the protection of the health and life of people moving in various mining areas.

#### **Acknowledgements**

**5. Conclusions**

*Mining Techniques - Past, Present and Future*

The determination of the deformation vectors from the differences between the adjusted geodetic point coordinate vectors obtained from at least two monitoring epochs is readily achievable if the original geometric and data structure of the geodetic network of the monitoring station between the individual monitoring epochs is strictly preserved. The presented research study provides both the theoretical and practical results of the possibility of solving deformation vectors in the geodetic network of the monitoring station in the case of its geometrical and data structure violation during the period of monitoring the movements of the earth surface, i.e., if the points of the geodetic network were damaged or completely destroyed between the different epochs of measurements, i.e., the geodetic network was nonhomogenous. The deformation vectors solved in accordance with the proposed specific theory of solving monitored deformations of the earth surface in the case of disturbance of the geodetic network homogeneity of the monitoring station provide via 3D models in GIS a reliable idea of spatial changes in the coordinates of the geodetic network points. The proposed theory of the specific deformation vector solution leads to a reliable support in investigation of various deformations of the earth surface, such as mining subsidence, landslides, geotectonic (recent move-

The largest values of the deformation vectors given in **Table 5** occurred at points No. 30 and No. 227. However, due to the fact that the deformation vectors tested at these points were not significant according to the test statistics, we can declare these points as static. The mentioned study case from the mining territory of the abandoned magnesite mine of Košice-Bankov confirmed the availability and applicability of the presented specific theory in the solution of the deformation vector in the deformation monitoring in the mining subsidence, where several violations of the geometric and data structure (homogeneity) of the geodetic network of the monitoring station occurred. Despite the validated method for the specific solution of the deformation vector in the geometric and data inhomogeneity of the geodetic network at the monitoring station, it should be stated that it would be preferable to maintain the homogeneity of the geodetic network during the whole monitoring period. Maintaining the homogeneity of the data of the geodetic network structure can ensure the permanent stabilization of the network points and the correct technically and physically implemented protection of the whole monitoring station from

3D model situations of the mining subsidence in GIS platform from the mining territory of Košice-Bankov were delivered to the municipality of the city of Košice (especially for the Department of Land Planning and Chief Architect of the City) to deal with a spatial planning for the future environmental reclaiming of this abandoned mining region, such as the magnesite mines of Košice-Bankov. The analysis of the deformation vectors at the geodetic network points of the monitoring station located in the mining subsidence and in the surrounding mining territory of the abandoned magnesite mine of Košice-Bankov was important in defining and specifying the subsidence edges and the subsidence zones with a number of dangerous cracks and fissures. The very precise identification of 3D position of such delimitation of the mining subsidence constituted the basic document for the plan of the municipality of the city of Košice for the reclaiming of the entire devastated mining landscape of Košice-Bankov. The revitalization of the Košice-Bankov recreational and tourist zone with the adjacent urban forest park has been achieved through the comprehensive reclaiming that devastated mining landscape. The variability of 3D models of the mining subsidence allowed a wide spectrum at modeling of natural and also industrial disasters in the former mining territory of Košice-Bankov.

ments), movements of dams, and other important building objects.

unexpected external interventions into such a station.

**150**

This research study was supported by the Slovak Research and Development Agency under the contract No. SK-CN-RD-18-0015 and by the European Union within the Interreg ENI CBC Programme under the contract No. HUSKROUA/ 1702/8.1/0065 and by the Scientific Grant Agency of the Ministry of Education, Science, Research and Sports of the Slovak Republic under the contract No. VEGA 1/0839/18.

#### **Author details**

Vladimír Sedlák Pavol Jozef Šafárik University in Košice, Slovakia

\*Address all correspondence to: vladimir.sedlak@upjs.sk

© 2020 The Author(s). Licensee IntechOpen. This chapter is distributed under the terms of the Creative Commons Attribution License (http://creativecommons.org/licenses/ by/3.0), which permits unrestricted use, distribution, and reproduction in any medium, provided the original work is properly cited.

### **References**

[1] Knothe S, editor. Forecasting the Influence of Mining (in Polish). Katowice: Śląsk Publishing House; 1984

[2] Whittaker BN, Reddish DJ, editors. Subsidence: Occurrence, Prediction and Control. 3rd ed. Amsterdam: Elsevier; 1989. p. 528

[3] Cui X, Miao X, Wang J, Yang S, Liu H, Hu X. Improved prediction of differential subsidence caused by underground mining. International Journal of Rock Mechanics and Mining Sciences. 2000;**37**(4):615-627. DOI: 10.1016/S1365-1609(99)00125-2

[4] Díaz-Fernández ME, Álvarez-Fernández MI, Álvarez-Vigil AE. Computation of influence functions for automatic mining subsidence prediction. Computational Geosciences. 2010;**14**(1):83-103. DOI: 10.1007/ s10596-009-9134-1

[5] Djamaluddin I, Mitani Z, Esaki T. Evaluation of ground movement and damage to structures from Chinese coal mining using a new GIS coupling model. International Journal of Rock Mechanics and Mining Sciences. 2011;**48**(3): 380-393. DOI: 10.1016/j.ijrmms.2011. 01.004

[6] Kratzsch H, editor. Mining Subsidence Engineering. Berlin: Springer-Verlag; 1983

[7] Sedlák V. Mathematical testing the edges of subsidence in undermined areas. Journal of Mining Science. 2014; **50**(3):465-474. DOI: 10.1134/ S1062739114030089

[8] Bauer RA. Mine subsidence in Illinois: Facts for homeowners. In: Circular 569 [Internet]. Champaign: Illinois State Geological Survey; 2006. p. 20. Available from: https://www2. illinois.gov/iema/Mitigation/Docume nts/Link\_Mine\_Subsidence\_Facts\_

Homeowners.pdf [Accessed: 24 September 2016]

[9] Colorado Geological Survey: Mine Subsidence [Internet]. 2016. Available from: http://coloradogeologicalsurvey. org/geologic-hazards/subsidence-mine/ [Accessed: 26 September 2016]

longwall coal mines, Slovakia. In: Borchers WJ, editor. Land Subsidence: Case Studies and Current Research. Belmont: U.S. Geological Survey; 1998.

*DOI: http://dx.doi.org/10.5772/intechopen.91461*

Slovakia. In: Proceedings of the 6th International Symposium on Land Subsidence–SISOLS 2000; 24–29 September; Ravenna. Ravenna: C.N.R.; 2005. Vol. II, 2000. pp. 139-150

[25] Sedlák V. Possibilities at modelling surface movements in GIS in the Košice depression, Slovakia. RMZ-Materials and Geoenvironment. 2004;**51**(4):

[26] Wright P, Stow R. Detecting mining subsidence from space. International Journal of Remote Sensing. 1999;**20**(6):

1183-1188. DOI: 10.1080/

[27] Koníček P, Souček K, Staš L, Singh R. Long-hole destress blasting for

[28] Strazalowski P, Scigala R. The example of linear discontinuous deformations caused by underground extraction. Transection of the VŠB - Technical University of Ostrava, Civil Engineering Series. 2005;**V**(2):193-198

[29] Li PX, Tan ZX, Deng KZ. Calculation of maximum ground movement and deformation caused by mining. Transactions of the Nonferrous Metals Society of China. 2011;**21**(3): s562-s569. DOI: 10.1016/S1003-6326

[30] Christensen R. General Gauss-Markov models. In: Christensen R, editor. Plane Answers to Complex Questions/the Theory of Linear Models. 4th ed. New York: Springer; 2011. pp. 237-266

[31] Gene H, Golub GH, Van Loan ChF, editors. Matrix Computations. 4th ed. Baltimore: JHU Press; 2013. p. 756

[32] Groß J. The general Gauss-Markov model with possibly singular dispersion

(12)61641-0

underground coal mining. International Journal of Rock Mechanics and Mining Sciences. 2013;**61**:141-153. DOI: 10.1016/j.ijrmms.2013.02.001

rockburst control during deep

014311699212939

2127-2133

*Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming…*

[18] Cai J, Wang J, Wu J, Hu C, Grafarend E, Chen J. Horizontal deformation rate analysis based on multiepoch GPS measurements in Shanghai. Journal of Surveying

Engineering. 2008;**134**(4):132-137. DOI: 10.1061/(ASCE)0733-9453(2008)134:4

[19] Can E, Mekik Ç, Kuşçu Ş, Akçın H. Computation of subsidence parameters resulting from layer movements postoperations of underground mining. Journal of Structural Geology. 2013;**47**: 16-24. DOI: 10.1016/ j.jsg.2012.11.005

[20] Hu LY. Gradual deformation and iterative calibration of Gaussian-related stochastic models. Mathematical Geology. 2000;**32**(1):87-108. DOI:

[21] Lü WC, Cheng SG, Yang HS, Liu DP. Application of GPS technology to build a mine-subsidence observation station. Journal of China University of Mining and Technology. 2008;**18**(3):377-380. DOI: 10.1016/S1006-1266(08)60079-6

[22] Marschalko M, Fuka M, Treslin L. Measurements by the method of precise inclinometry on locality affected by mining activity. Archives of Mining Sciences. 2008;**53**(3):397-414

[23] Ng AH, Ge L, Zhang K, Chang HC, Li X, Rizos C, et al. Deformation mapping in three dimensions for underground mining using InSAR - Southern highland coalfield in New South Wales, Australia. International Journal of Remote Sensing. 2011;**32**(22):

7227-7256. DOI: 10.1080/ 01431161.2010.519741

**153**

[24] Sedlák V. GPS measurement of geo-tectonic recent movements in east

10.1023/A:1007506918588

pp. 257-263

(132)

[10] Donnelly LJ, Reddish DJ. The development of surface steps during mining subsidence: "Not due to fault reactivation". Engineering Geology. 1994;**36**(3–4):243-255. DOI: 10.1016/ 0013-7952(94)90006-X

[11] Sedlák V, editor. Modelling Subsidence Development at the Mining Damages. Košice: Štroffek; 1997. p. 52

[12] Sedlák V. Modelling subsidence deformations at the Slovak coalfields. Kuwait Journal of Science & Engineering. 1997;**24**(2):339-349

[13] Alehossein H. Back of envelope mining subsidence estimation. Australian Geomechanics. 2009;**44**(1):29-32

[14] Can E, Mekik Ç, Kuşçu Ş, Akçın H. Monitoring deformations on engineering structures in Kozlu Hard Coal Basin. Natural Hazards. 2013;**65**(3):2311-2330. DOI: 10.1007/s11069-012-0477-x

[15] Jung HC, Kim SW, Jung HS, Min KD, Won JS. Satellite observation of coal mining subsidence by persistent scatterer analysis. Engineering Geology. 2007;**92**(1):1-13. DOI: 10.1016/j.enggeo. 2007.02.007

[16] Sedlák V, Kunák L, Havlice K, Šadera M. Modelling deformations in land subsidence development at the Slovak coalfields. Survey Ireland. 1995;**12**(13): 25-29

[17] Sedlák V. Measurement and prediction of land subsidence above *Specific Solution of Deformation Vector in Land Subsidence for GIS Applications to Reclaiming… DOI: http://dx.doi.org/10.5772/intechopen.91461*

longwall coal mines, Slovakia. In: Borchers WJ, editor. Land Subsidence: Case Studies and Current Research. Belmont: U.S. Geological Survey; 1998. pp. 257-263

**References**

1989. p. 528

[1] Knothe S, editor. Forecasting the Influence of Mining (in Polish).

Katowice: Śląsk Publishing House; 1984

*Mining Techniques - Past, Present and Future*

Homeowners.pdf [Accessed:

[9] Colorado Geological Survey: Mine Subsidence [Internet]. 2016. Available from: http://coloradogeologicalsurvey. org/geologic-hazards/subsidence-mine/

[Accessed: 26 September 2016]

0013-7952(94)90006-X

[11] Sedlák V, editor. Modelling

Subsidence Development at the Mining Damages. Košice: Štroffek; 1997. p. 52

[12] Sedlák V. Modelling subsidence deformations at the Slovak coalfields.

[13] Alehossein H. Back of envelope mining subsidence estimation. Australian

Geomechanics. 2009;**44**(1):29-32

[15] Jung HC, Kim SW, Jung HS, Min KD, Won JS. Satellite observation of coal mining subsidence by persistent scatterer analysis. Engineering Geology. 2007;**92**(1):1-13. DOI: 10.1016/j.enggeo.

[16] Sedlák V, Kunák L, Havlice K, Šadera M. Modelling deformations in land subsidence development at the Slovak coalfields. Survey Ireland. 1995;**12**(13):

[17] Sedlák V. Measurement and prediction of land subsidence above

2007.02.007

25-29

[14] Can E, Mekik Ç, Kuşçu Ş, Akçın H. Monitoring deformations on engineering structures in Kozlu Hard Coal Basin. Natural Hazards. 2013;**65**(3):2311-2330. DOI: 10.1007/s11069-012-0477-x

Kuwait Journal of Science & Engineering. 1997;**24**(2):339-349

[10] Donnelly LJ, Reddish DJ. The development of surface steps during mining subsidence: "Not due to fault reactivation". Engineering Geology. 1994;**36**(3–4):243-255. DOI: 10.1016/

24 September 2016]

[2] Whittaker BN, Reddish DJ, editors. Subsidence: Occurrence, Prediction and Control. 3rd ed. Amsterdam: Elsevier;

[3] Cui X, Miao X, Wang J, Yang S, Liu H, Hu X. Improved prediction of differential subsidence caused by underground mining. International Journal of Rock Mechanics and Mining Sciences. 2000;**37**(4):615-627. DOI: 10.1016/S1365-1609(99)00125-2

[4] Díaz-Fernández ME, Álvarez-Fernández MI, Álvarez-Vigil AE. Computation of influence functions for

prediction. Computational Geosciences. 2010;**14**(1):83-103. DOI: 10.1007/

[5] Djamaluddin I, Mitani Z, Esaki T. Evaluation of ground movement and damage to structures from Chinese coal mining using a new GIS coupling model. International Journal of Rock Mechanics and Mining Sciences. 2011;**48**(3): 380-393. DOI: 10.1016/j.ijrmms.2011.

automatic mining subsidence

[6] Kratzsch H, editor. Mining Subsidence Engineering. Berlin:

**50**(3):465-474. DOI: 10.1134/

[8] Bauer RA. Mine subsidence in Illinois: Facts for homeowners. In: Circular 569 [Internet]. Champaign: Illinois State Geological Survey; 2006. p. 20. Available from: https://www2. illinois.gov/iema/Mitigation/Docume nts/Link\_Mine\_Subsidence\_Facts\_

[7] Sedlák V. Mathematical testing the edges of subsidence in undermined areas. Journal of Mining Science. 2014;

Springer-Verlag; 1983

S1062739114030089

s10596-009-9134-1

01.004

**152**

[18] Cai J, Wang J, Wu J, Hu C, Grafarend E, Chen J. Horizontal deformation rate analysis based on multiepoch GPS measurements in Shanghai. Journal of Surveying Engineering. 2008;**134**(4):132-137. DOI: 10.1061/(ASCE)0733-9453(2008)134:4 (132)

[19] Can E, Mekik Ç, Kuşçu Ş, Akçın H. Computation of subsidence parameters resulting from layer movements postoperations of underground mining. Journal of Structural Geology. 2013;**47**: 16-24. DOI: 10.1016/ j.jsg.2012.11.005

[20] Hu LY. Gradual deformation and iterative calibration of Gaussian-related stochastic models. Mathematical Geology. 2000;**32**(1):87-108. DOI: 10.1023/A:1007506918588

[21] Lü WC, Cheng SG, Yang HS, Liu DP. Application of GPS technology to build a mine-subsidence observation station. Journal of China University of Mining and Technology. 2008;**18**(3):377-380. DOI: 10.1016/S1006-1266(08)60079-6

[22] Marschalko M, Fuka M, Treslin L. Measurements by the method of precise inclinometry on locality affected by mining activity. Archives of Mining Sciences. 2008;**53**(3):397-414

[23] Ng AH, Ge L, Zhang K, Chang HC, Li X, Rizos C, et al. Deformation mapping in three dimensions for underground mining using InSAR - Southern highland coalfield in New South Wales, Australia. International Journal of Remote Sensing. 2011;**32**(22): 7227-7256. DOI: 10.1080/ 01431161.2010.519741

[24] Sedlák V. GPS measurement of geo-tectonic recent movements in east Slovakia. In: Proceedings of the 6th International Symposium on Land Subsidence–SISOLS 2000; 24–29 September; Ravenna. Ravenna: C.N.R.; 2005. Vol. II, 2000. pp. 139-150

[25] Sedlák V. Possibilities at modelling surface movements in GIS in the Košice depression, Slovakia. RMZ-Materials and Geoenvironment. 2004;**51**(4): 2127-2133

[26] Wright P, Stow R. Detecting mining subsidence from space. International Journal of Remote Sensing. 1999;**20**(6): 1183-1188. DOI: 10.1080/ 014311699212939

[27] Koníček P, Souček K, Staš L, Singh R. Long-hole destress blasting for rockburst control during deep underground coal mining. International Journal of Rock Mechanics and Mining Sciences. 2013;**61**:141-153. DOI: 10.1016/j.ijrmms.2013.02.001

[28] Strazalowski P, Scigala R. The example of linear discontinuous deformations caused by underground extraction. Transection of the VŠB - Technical University of Ostrava, Civil Engineering Series. 2005;**V**(2):193-198

[29] Li PX, Tan ZX, Deng KZ. Calculation of maximum ground movement and deformation caused by mining. Transactions of the Nonferrous Metals Society of China. 2011;**21**(3): s562-s569. DOI: 10.1016/S1003-6326 (12)61641-0

[30] Christensen R. General Gauss-Markov models. In: Christensen R, editor. Plane Answers to Complex Questions/the Theory of Linear Models. 4th ed. New York: Springer; 2011. pp. 237-266

[31] Gene H, Golub GH, Van Loan ChF, editors. Matrix Computations. 4th ed. Baltimore: JHU Press; 2013. p. 756

[32] Groß J. The general Gauss-Markov model with possibly singular dispersion matrix. Statistical Papers. 2004;**45**(3): 311-336. DOI: 10.1007/BF02777575

[33] Lindgren F, Ruel H, Lindström J. An explicit link between Gaussian fields and Gaussian Markov random fields: The stochastic partial differential equation approach. Journal of the Royal Statistical Society, Series B: Statistical Methodology. 2011;**73**(4):423-498. DOI: 10.1111/j.1467-9868.2011.00777.x

[34] Lehmann EL, Romano JP, editors. Testing Statistical Hypotheses. 3rd ed. New York: Springer; 2008. p. 786

[35] Blachowski J. Application of GIS spatial regression methods in assessment of land subsidence in complicated mining conditions: Case study of the Walbrzych coal mine (SW Poland). Natural Hazards. 2016;**84**(2):997-1014. DOI: 10.1007/s11069-016-2470-2

[36] Gallay M, Kaňuk J, Hochmuth Z, Meneely JD, Hofierka J, Sedlák V. Large-scale and high-resolution 3-D cave mapping by terrestrial laser scanning: A case study of the Domica Cave, Slovakia. International Journal of Speleology. 2015;**44**(3):277-291. DOI: 10.5038/1827-806X.44.3.6

[37] Kaňuk J, Gallay M, Hofierka J. Generating time series of virtual 3-D city models using a retrospective approach. Landscape and Urban Planning. 2015;**139**:40-53. DOI: 10.1016/ j.landurbplan.2015.02.015

[38] Yang KM, Xiao JB, Duan MT, Pang B, Wang YB, Wang R. Geodeformation information extraction and GIS analysis on important buildings by underground mining subsidence. In: Proceedings of the International Conference on Information Engineering and Computer Science–ICIECS 2009. 19–20 December, 2009; Wuhan. Wuhan: IEEE; 2009. p. 4

[39] Yang KM, Ma JT, Pang B, Wang YB, Wang R, Duan MT. 3D visual technology of geo-deformation disasters induced by mining subsidence based on ArcGIS engine. Key Engineering Materials. 2012; **500**:428-436. DOI: 10.4028/www. scientific.net/KEM.500.428

**155**

**Chapter 9**

*Xiaohan Yang*

**Abstract**

and safe mining.

rock mechanics

**1. Introduction**

Energy Aspect

Coal Burst: A State of the Art on

Mechanism and Prevention from

Coal burst continues to be one of the most catastrophic safety hazards faced by future mining as the stress environment will be more complicated with the increase of mining depth. Many chief coal mining countries including Poland, Czech Republic, the U.S., China, and Australia have experienced fatal accidents caused by coal burst and conducted comprehensive research on the driving forces and solving technologies related to coal burst. In this chapter, the research outcomes of the mechanism, risk evaluation, risk monitoring, and prevention of coal burst are reviewed, which is helpful for mining researchers and engineers to understand and control the safety hazards caused by coal burst, and, hence, to achieve sustainable

**Keywords:** coal burst, underground mining, mining safety, dynamic hazards,

Coal burst, which refers to the violent and catastrophic failure of coal, is a serious safety hazard for underground coal mines, and it has attracted intensive research interests from mining and geological scholars [1]. In 1738, the first

recorded coal burst took place in England [2, 3]. Since then, both the frequency and severity of coal burst increased with mining depth [2, 4, 5]. As shown in **Table 1**, coal burst has been a serious security issue that many countries face for decades. Coal burst has been recognized as a serious risk for Australia's underground coal mines following a fatal coal burst accident at the Austar Coal Mine [6, 7]. Because of lacking coal burst experience, it is difficult to find mature theories and technologies in Australian to explain, predict, monitor, or control coal burst. It is an urgent task to develop a coal burst risk assessment methodology and prevention technology for Australian coal mines. Extensive study has been conducted around the mechanism, prediction, and prevention of coal burst [5] by scholars around the world. Some necessary conditions of coal burst such as stiffness, dynamic load, and mechanical

In terms of energy, coal burst is the energy accumulation and releasing process of a coal body. Coal burst monitoring, such as acoustic emission, electromagnetic radiation, micro-seismic, infrared, and other methods, is the monitoring of different energy forms released during coal burst [8, 9]. The cause of the coal ejection and roadway destruction is the elastic energy stored in the coal [10]. Therefore,

property are found based on previous decades' research.

#### **Chapter 9**

matrix. Statistical Papers. 2004;**45**(3): 311-336. DOI: 10.1007/BF02777575

*Mining Techniques - Past, Present and Future*

of geo-deformation disasters induced by mining subsidence based on ArcGIS engine. Key Engineering Materials. 2012; **500**:428-436. DOI: 10.4028/www. scientific.net/KEM.500.428

[33] Lindgren F, Ruel H, Lindström J. An explicit link between Gaussian fields and Gaussian Markov random fields: The stochastic partial differential equation approach. Journal of the Royal Statistical Society, Series B: Statistical Methodology. 2011;**73**(4):423-498. DOI: 10.1111/j.1467-9868.2011.00777.x

[34] Lehmann EL, Romano JP, editors. Testing Statistical Hypotheses. 3rd ed. New York: Springer; 2008. p. 786

[35] Blachowski J. Application of GIS spatial regression methods in assessment of land subsidence in complicated mining conditions: Case study of the Walbrzych coal mine (SW Poland). Natural Hazards. 2016;**84**(2):997-1014. DOI: 10.1007/s11069-016-2470-2

[36] Gallay M, Kaňuk J, Hochmuth Z, Meneely JD, Hofierka J, Sedlák V. Large-scale and high-resolution 3-D cave mapping by terrestrial laser scanning: A case study of the Domica Cave, Slovakia. International Journal of Speleology. 2015;**44**(3):277-291. DOI:

10.5038/1827-806X.44.3.6

j.landurbplan.2015.02.015

Wuhan: IEEE; 2009. p. 4

**154**

[37] Kaňuk J, Gallay M, Hofierka J. Generating time series of virtual 3-D city models using a retrospective approach. Landscape and Urban

[38] Yang KM, Xiao JB, Duan MT, Pang B, Wang YB, Wang R. Geodeformation information extraction and GIS analysis on important buildings by underground mining subsidence. In: Proceedings of the International

Planning. 2015;**139**:40-53. DOI: 10.1016/

Conference on Information Engineering and Computer Science–ICIECS 2009. 19–20 December, 2009; Wuhan.

[39] Yang KM, Ma JT, Pang B, Wang YB, Wang R, Duan MT. 3D visual technology

## Coal Burst: A State of the Art on Mechanism and Prevention from Energy Aspect

*Xiaohan Yang*

#### **Abstract**

Coal burst continues to be one of the most catastrophic safety hazards faced by future mining as the stress environment will be more complicated with the increase of mining depth. Many chief coal mining countries including Poland, Czech Republic, the U.S., China, and Australia have experienced fatal accidents caused by coal burst and conducted comprehensive research on the driving forces and solving technologies related to coal burst. In this chapter, the research outcomes of the mechanism, risk evaluation, risk monitoring, and prevention of coal burst are reviewed, which is helpful for mining researchers and engineers to understand and control the safety hazards caused by coal burst, and, hence, to achieve sustainable and safe mining.

**Keywords:** coal burst, underground mining, mining safety, dynamic hazards, rock mechanics

#### **1. Introduction**

Coal burst, which refers to the violent and catastrophic failure of coal, is a serious safety hazard for underground coal mines, and it has attracted intensive research interests from mining and geological scholars [1]. In 1738, the first recorded coal burst took place in England [2, 3]. Since then, both the frequency and severity of coal burst increased with mining depth [2, 4, 5]. As shown in **Table 1**, coal burst has been a serious security issue that many countries face for decades. Coal burst has been recognized as a serious risk for Australia's underground coal mines following a fatal coal burst accident at the Austar Coal Mine [6, 7]. Because of lacking coal burst experience, it is difficult to find mature theories and technologies in Australian to explain, predict, monitor, or control coal burst. It is an urgent task to develop a coal burst risk assessment methodology and prevention technology for Australian coal mines. Extensive study has been conducted around the mechanism, prediction, and prevention of coal burst [5] by scholars around the world. Some necessary conditions of coal burst such as stiffness, dynamic load, and mechanical property are found based on previous decades' research.

In terms of energy, coal burst is the energy accumulation and releasing process of a coal body. Coal burst monitoring, such as acoustic emission, electromagnetic radiation, micro-seismic, infrared, and other methods, is the monitoring of different energy forms released during coal burst [8, 9]. The cause of the coal ejection and roadway destruction is the elastic energy stored in the coal [10]. Therefore,


#### **Table 1.**

*Coal burst occurrence and fatalities by country/region [7].*

it is significant to have an understanding of energy release mode in the coal burst process, especially the magnitude of coal burst energy. Coal burst is regarded as a dynamic disaster since it is shown in many studies that coal burst is closely related to dynamic load [11]. It is believed that hard rock is more prone to violent failure than soft rock [12]. Due to the difference in physical and mechanical properties, different coal seams have a different coal burst propensity. Therefore, changing coal mechanical property is a promising method for coal burst mitigation. Water infusion can mitigate coal burst propensity through increasing moisture content of coal [13]. In this chapter, the coal burst driving forces, solving techniques, and monitoring methods are reviewed from energy aspects.

#### **2. Potential driving factors**

#### **2.1 Mining depth**

Mining depth has been identified as an important factor for the formation of coal burst. According to the analysis of coal burst cases in Poland and China, LM Dou found that the first coal burst accident in coal mines generally happened when mining depth approached 350 m and the frequency and severity of coal burst sharply increases with mining depth changing from 350 to 600 m [14]. Iannacchione and Zelanko found that nearly all coal bursts in the main coal fields of the U.S. occurred at depths greater than 300 meters, and most were at depths exceeding 400 m [15]. The contribution of mining depth to coal burst mainly results from the increasing gravitational stress. More strain energy will be stored in the coal under high gravitational stress condition [16]. Besides, for coal mines in China and the U.S., hard sandstone roof seems the common geological feature for deep mining, which can further result in a large accumulation of energy or a catastrophic dynamic load [17, 18]. The potential influence of hard roof (roof stiffness) also will be discussed in another section of this paper. The mining depths of two coal mines with coal burst accident in Australia are both around 500 m [19]. Hence, the strain energy accumulation led by high gravitational stress plays an important role in the formation of coal burst accidents that happened in Australia as the mining depth of these coal mines is already beyond the mining depth of majority of burst accidents revealed by international research.

More seriously, almost all coal mines in Australia have plans for deeper mining, which means the stress environments will be more complicated and more energy will be stored in coal seams [20].

**157**

**Figure 1.**

*Stress concentration caused by geological structures.*

*Coal Burst: A State of the Art on Mechanism and Prevention from Energy Aspect*

It has been shown by numerous studies that the complicated geological structures caused by folds, faults, and coal seam thickness variation have a noticeable influence on the coal burst occurrence [21]. Dou et al. found that 72% coal burst accidents in Longfeng Colliery were related to faults [16]. The numerical study conducted by Chen et al. found that stress will concentrate near the coal face when the coal face approaches fault [22]. Mark found that coal burst accidents in the U.S. have a close relationship with faults [23]. Folds, which are created by compressional tectonic stress, may have high residual tectonic stress in the geological structures. Through the stress regression analysis of Huanghuiyan Colliery, Jiang et al. found that stress concentration tends to exit at the area near syncline axis [24]. The influence of geological structures on stress distribution is shown as **Figure 1**.

Stiffness of the surrounding rocks is one of the main factors giving rise to coal burst. Bieniawski found that rock samples are more prone to violent failure under the loading machine with high stiffness. The uniaxial compression tests of sample composed by coal and rock found that most elastic energy is stored in the coal part of the compound sample and the burst potential of the sample is positively related to the thickness of the rock part [25, 26]. Through theoretical analysis, Yang found that energy will flow from high stiffness material to low stiffness material [17]. Hence, the high stiffness of surrounding rocks will enhance the energy accumulation in coal seam. In addition, as shown in **Figure 2**, the strength of coal tends to have rapidly decreased under the high stiffness environment [27]. Generally, the high stiffness environment is related to the heavy and hard sandstone layer above the coal seam [28]. Sometimes, the thickness of sandstone layer can reach tens or

*DOI: http://dx.doi.org/10.5772/intechopen.91988*

**2.2 Geological structures**

**2.3 Surrounding rocks' stiffness**

even hundreds of meters [16].

*Coal Burst: A State of the Art on Mechanism and Prevention from Energy Aspect DOI: http://dx.doi.org/10.5772/intechopen.91988*

#### **2.2 Geological structures**

*Mining Techniques - Past, Present and Future*

**Country/region Time** 

*Coal burst occurrence and fatalities by country/region [7].*

Czech Republic/ Poland

**Table 1.**

**period**

USA 1943–2003 – 78

it is significant to have an understanding of energy release mode in the coal burst process, especially the magnitude of coal burst energy. Coal burst is regarded as a dynamic disaster since it is shown in many studies that coal burst is closely related to dynamic load [11]. It is believed that hard rock is more prone to violent failure than soft rock [12]. Due to the difference in physical and mechanical properties, different coal seams have a different coal burst propensity. Therefore, changing coal mechanical property is a promising method for coal burst mitigation. Water infusion can mitigate coal burst propensity through increasing moisture content of coal [13]. In this chapter, the coal burst driving forces, solving techniques, and

**Number of coal bursts**

1983–2003 190 122

Ruhr, Germany 1973–1992 50 27 [4]

USA 1983–2013 337 20 [21] Mainland China 1933–1996 4000 400 [5] Mainland China 2006–2013 >35 >300 [36]

**Number of fatalities**

**Reference**

Mining depth has been identified as an important factor for the formation of coal burst. According to the analysis of coal burst cases in Poland and China, LM Dou found that the first coal burst accident in coal mines generally happened when mining depth approached 350 m and the frequency and severity of coal burst sharply increases with mining depth changing from 350 to 600 m [14]. Iannacchione and Zelanko found that nearly all coal bursts in the main coal fields of the U.S. occurred at depths greater than 300 meters, and most were at depths exceeding 400 m [15]. The contribution of mining depth to coal burst mainly results from the increasing gravitational stress. More strain energy will be stored in the coal under high gravitational stress condition [16]. Besides, for coal mines in China and the U.S., hard sandstone roof seems the common geological feature for deep mining, which can further result in a large accumulation of energy or a catastrophic dynamic load [17, 18]. The potential influence of hard roof (roof stiffness) also will be discussed in another section of this paper. The mining depths of two coal mines with coal burst accident in Australia are both around 500 m [19]. Hence, the strain energy accumulation led by high gravitational stress plays an important role in the formation of coal burst accidents that happened in Australia as the mining depth of these coal mines is already beyond the mining depth of majority of burst accidents revealed by international research. More seriously, almost all coal mines in Australia have plans for deeper mining, which means the stress environments will be more complicated and more energy

monitoring methods are reviewed from energy aspects.

**2. Potential driving factors**

will be stored in coal seams [20].

**2.1 Mining depth**

**156**

It has been shown by numerous studies that the complicated geological structures caused by folds, faults, and coal seam thickness variation have a noticeable influence on the coal burst occurrence [21]. Dou et al. found that 72% coal burst accidents in Longfeng Colliery were related to faults [16]. The numerical study conducted by Chen et al. found that stress will concentrate near the coal face when the coal face approaches fault [22]. Mark found that coal burst accidents in the U.S. have a close relationship with faults [23]. Folds, which are created by compressional tectonic stress, may have high residual tectonic stress in the geological structures. Through the stress regression analysis of Huanghuiyan Colliery, Jiang et al. found that stress concentration tends to exit at the area near syncline axis [24]. The influence of geological structures on stress distribution is shown as **Figure 1**.

#### **2.3 Surrounding rocks' stiffness**

Stiffness of the surrounding rocks is one of the main factors giving rise to coal burst. Bieniawski found that rock samples are more prone to violent failure under the loading machine with high stiffness. The uniaxial compression tests of sample composed by coal and rock found that most elastic energy is stored in the coal part of the compound sample and the burst potential of the sample is positively related to the thickness of the rock part [25, 26]. Through theoretical analysis, Yang found that energy will flow from high stiffness material to low stiffness material [17]. Hence, the high stiffness of surrounding rocks will enhance the energy accumulation in coal seam. In addition, as shown in **Figure 2**, the strength of coal tends to have rapidly decreased under the high stiffness environment [27]. Generally, the high stiffness environment is related to the heavy and hard sandstone layer above the coal seam [28]. Sometimes, the thickness of sandstone layer can reach tens or even hundreds of meters [16].

**Figure 1.** *Stress concentration caused by geological structures.*

**Figure 2.**

*Effect of stiffness of the loading system on the behavior of coal failure [17].*

#### **2.4 Micro-seismicity**

Micro-seismicity refers to the regional small-scale seismic events that are undetectable by earthquake monitoring stations due to their small-scale energy compared with earthquakes. However, for underground coal mines, the energy released by micro-seismicity also is an important energy source for coal burst formation. Intensive micro-seismicity has been observed in most coal mines with high bursts risk in Poland, China, and the U.S. [29–31]. Micro-seismicity can be detected and located by specific micro-seismic monitoring apparatus. Deep research has been made by many researchers on the monitoring of dynamic load and identifying high burst potential areas through micro-seismic monitoring [32–34].

#### **3. Previous mechanism**

The study of the coal burst mechanism aims to explain the causes of coal burst from two perspectives: force source and coal's physical properties. As a type of coal failure, coal burst should meet the conditions of coal failure. That is, the stress loaded on coal exceeds the strength of coal when coal burst occurs, which is named strength theory by some scholars [16]. Satisfying strength theory is one of the conditions required by coal burst. Under static loading condition, coal burst does not always happen when the ultimate strength is reached. It has been pointed out that coal strength will change under dynamic load. Research has shown that the coal failure behavior is affected by loading rate as well [35]. In the actual situation, the strength theory of coal burst becomes more complex as the coal body is under the collective effect of static load (overburden weight) and dynamic load. Dou et al. [14] studied the dynamic load required by coal burst at different static load levels. Through a series of follow-up studies, LM Dou put forward the dynamic and static load superposition theory of coal burst [36, 37]. The strength theory of coal burst

**159**

**4. Prevention methods**

**4.1 Evaluation**

*Coal Burst: A State of the Art on Mechanism and Prevention from Energy Aspect*

under dynamic load should be based on the dynamic strength of coal. Cook found that marble only has violent failure when the stiffness of the test machine is greater than the stiffness of the specimen [2, 16]. The compressive experiment of samples composed of coal and rock showed that violent failure always occurred in the layer with minimum stiffness [25, 38]. That is, the necessary condition for coal burst of a pillar or rib is that the stiffness of the roof and floor is greater than that of the coal seam. In most cases, the stiffness of coal seam is minimal relative to roof and floor.

*Schematic diagram of coal burst propensity index [40]. (a) Determination of WET and (b) determination of KE.*

It is found that the post-failure curve of hard rock is steeper than that of soft rock. This means that hard rock is more likely to fail instantaneously. Bieniawski et al. [39] believe that hard rock is much more prone to violent rupture than soft rock. It is necessary to explain that the hard rock and soft rock here are classified in terms of strength. Bieniawski proposed two indices, elastic strain energy index (WET) and bursting energy index (KE), to measure the rock burst tendency of different rocks. As shown in **Figure 1**, elastic strain energy index is the ratio between elastic energy (Ee) and plastic energy (Ep) when the specimen is loaded to at least 80% of the strength and then unloaded [2]. KE is the ratio between Eb and Ea [2]. Eb represents the energy storage before strength while Ea means deformation energy consumed after the peak value. It is proved by in suit and experimental data that coal with high WET and KE value has a high tendency for violent failure [2, 4, 25]. These two indices describe the proportion of elastic energy during coal burst. Different rock types have different burst tendency and different energy storage and releasing behavior. Due to the difference in physical and mechanical properties, the WET and KE values of different coal seams vary widely as well. Theoretically speaking, coal has no burst ability when the WET and KE values are low enough. The ability or property of coal burst is called coal burst propensity by Chinese scholars. Four indices including WET and KE are summarized as coal burst propensity indices by Chinese scholars and have become a good indicator of coal burst risk of different coal seams. Coal burst propensity index describes the proportions of different energies. The successful application of the coal burst propensity index method indicates that elastic energy and coal burst are closely related. Coal has the ability to store and

That is, coal failure in coal mines generally meet stiffness conditions.

instantly release elastic energy in the premise of coal burst (**Figure 3**).

Based on the analysis of stress-strain curve of coal specimens under uniaxial compression stress, several special indices are published by different researchers to evaluate coal burst propensity. Russian and Poland coal mines adopt elastic strain energy index and bursting energy index to evaluate coal burst propensity [2, 4]. Zhang et al.

*DOI: http://dx.doi.org/10.5772/intechopen.91988*

**Figure 3.**

*Coal Burst: A State of the Art on Mechanism and Prevention from Energy Aspect DOI: http://dx.doi.org/10.5772/intechopen.91988*

**Figure 3.**

*Mining Techniques - Past, Present and Future*

**2.4 Micro-seismicity**

**Figure 2.**

monitoring [32–34].

**3. Previous mechanism**

Micro-seismicity refers to the regional small-scale seismic events that are undetectable by earthquake monitoring stations due to their small-scale energy compared with earthquakes. However, for underground coal mines, the energy released by micro-seismicity also is an important energy source for coal burst formation. Intensive micro-seismicity has been observed in most coal mines with high bursts risk in Poland, China, and the U.S. [29–31]. Micro-seismicity can be detected and located by specific micro-seismic monitoring apparatus. Deep research has been made by many researchers on the monitoring of dynamic load and identifying high burst potential areas through micro-seismic

*Effect of stiffness of the loading system on the behavior of coal failure [17].*

The study of the coal burst mechanism aims to explain the causes of coal burst from two perspectives: force source and coal's physical properties. As a type of coal failure, coal burst should meet the conditions of coal failure. That is, the stress loaded on coal exceeds the strength of coal when coal burst occurs, which is named strength theory by some scholars [16]. Satisfying strength theory is one of the conditions required by coal burst. Under static loading condition, coal burst does not always happen when the ultimate strength is reached. It has been pointed out that coal strength will change under dynamic load. Research has shown that the coal failure behavior is affected by loading rate as well [35]. In the actual situation, the strength theory of coal burst becomes more complex as the coal body is under the collective effect of static load (overburden weight) and dynamic load. Dou et al. [14] studied the dynamic load required by coal burst at different static load levels. Through a series of follow-up studies, LM Dou put forward the dynamic and static load superposition theory of coal burst [36, 37]. The strength theory of coal burst

**158**

*Schematic diagram of coal burst propensity index [40]. (a) Determination of WET and (b) determination of KE.*

under dynamic load should be based on the dynamic strength of coal. Cook found that marble only has violent failure when the stiffness of the test machine is greater than the stiffness of the specimen [2, 16]. The compressive experiment of samples composed of coal and rock showed that violent failure always occurred in the layer with minimum stiffness [25, 38]. That is, the necessary condition for coal burst of a pillar or rib is that the stiffness of the roof and floor is greater than that of the coal seam. In most cases, the stiffness of coal seam is minimal relative to roof and floor. That is, coal failure in coal mines generally meet stiffness conditions.

It is found that the post-failure curve of hard rock is steeper than that of soft rock. This means that hard rock is more likely to fail instantaneously. Bieniawski et al. [39] believe that hard rock is much more prone to violent rupture than soft rock. It is necessary to explain that the hard rock and soft rock here are classified in terms of strength. Bieniawski proposed two indices, elastic strain energy index (WET) and bursting energy index (KE), to measure the rock burst tendency of different rocks. As shown in **Figure 1**, elastic strain energy index is the ratio between elastic energy (Ee) and plastic energy (Ep) when the specimen is loaded to at least 80% of the strength and then unloaded [2]. KE is the ratio between Eb and Ea [2]. Eb represents the energy storage before strength while Ea means deformation energy consumed after the peak value. It is proved by in suit and experimental data that coal with high WET and KE value has a high tendency for violent failure [2, 4, 25]. These two indices describe the proportion of elastic energy during coal burst. Different rock types have different burst tendency and different energy storage and releasing behavior. Due to the difference in physical and mechanical properties, the WET and KE values of different coal seams vary widely as well. Theoretically speaking, coal has no burst ability when the WET and KE values are low enough. The ability or property of coal burst is called coal burst propensity by Chinese scholars. Four indices including WET and KE are summarized as coal burst propensity indices by Chinese scholars and have become a good indicator of coal burst risk of different coal seams. Coal burst propensity index describes the proportions of different energies. The successful application of the coal burst propensity index method indicates that elastic energy and coal burst are closely related. Coal has the ability to store and instantly release elastic energy in the premise of coal burst (**Figure 3**).

#### **4. Prevention methods**

#### **4.1 Evaluation**

Based on the analysis of stress-strain curve of coal specimens under uniaxial compression stress, several special indices are published by different researchers to evaluate coal burst propensity. Russian and Poland coal mines adopt elastic strain energy index and bursting energy index to evaluate coal burst propensity [2, 4]. Zhang et al. believe that the duration of failure process is the comprehensive reflection of energy accumulation and dissipation characteristics of coal [41]. They propose a dynamic failure time to evaluate coal burst propensity. Based on the correlation analysis of mass data, Qi et al. conclude that uniaxial compression strength of coal is a proper index of coal burst propensity evaluation as well [42]. In 2010, the China Coal Industry Association summarized these four indices as bursting liability indices of coal and published the standard test method of these four indices. Some researchers adopt these four indices to evaluate the burst propensity of rocks as well. It is has been proved by Russian, Poland, and Chinese experience that these four indices are good indicators to define the burst risk of coal seam. Besides, LM Dou et al. combined geological conditions and technical settings of mining together and proposed comprehensive index method based on the coal burst research in China [16].

#### **4.2 Monitoring**

Minimizing the safety risk caused by failure of instability rock/coal is an urgent and essential task for underground mines. Similar with the instantaneous failure of other brittle materials such as rock, concrete, and metal, the coal burst process is always associated with the release of rich geophysical signals including acoustic emission (AE) [43], micro-seismic [32] and electromagnetic radiation [44]. It is demonstrated by decades of research and in-field application that micro-seismic monitoring technology has a promising ability to locate potentially violent rock failure. Micro-seismic monitoring is a passive observation of very small-scale earthquakes that occur in the underground as a result of human activities such as mining, hydraulic fracturing, and underground gas storage. The phenomenon that stressed rock can release micro-level signal was discovered by two researchers of U.S. Bureau of Mines, Obert and Duvall, in 1938 [32, 34]. In the early 1960s, South African researchers developed a 16-channel micro-seismic system with positioning function for rock burst monitoring in gold mines [34]. In 1970, under the sponsorship of the U.S. Bureau of Mines, the Pennsylvania State Rock Mechanics Laboratory conducted a research project to investigate the application of micro-seismic techniques to coal mine safety [45]. Through decades' study of underground micro-seismic for mining operation, micro-seismic system has been a basic and valuable monitoring tool for metal and coal mines worldwide. It provides a continuous and real-time 4D (three dimension location and time) record of seismicity associated with rock failure in the monitoring region.

#### **4.3 Controlling**

The widely used coal burst controlling methods include provocative blasting, long-term water infusion, hydro-fracturing, de-stress drilling, and protective seam mining [46]. Dou et al. proposed the intensity weakening theory to guide the coal burst control from the aspect of energy [16]. Based on the energy aspects, the key to coal burst prevention are: (1) softening coal by changing the physical and mechanical properties of coal. The burst tendency or burst scale of soft coal will be mitigated as the energy storage ability of coal has been reduced. The main methods of coal body softening are blasting and water infusion. (2) Transferring stress to deep regions and reducing the stress level of coal, which can reduce energy storage as well. The main methods are pressure relief blasting, roof pre-splitting blasting, roof cutting blasting, protection seam mining, hydraulic roof fracturing, and large diameter pressure relief drilling. (3) Releasing energy by artificially induced coal burst under low stress level. The main methods are pressure relief blasting and large diameter pressure relief drilling.

**161**

**Author details**

Xiaohan Yang

NSW, Australia

*Coal Burst: A State of the Art on Mechanism and Prevention from Energy Aspect*

School of Civil, Mining and Environmental Engineering, University of Wollongong,

© 2020 The Author(s). Licensee IntechOpen. This chapter is distributed under the terms of the Creative Commons Attribution License (http://creativecommons.org/licenses/ by/3.0), which permits unrestricted use, distribution, and reproduction in any medium,

\*Address all correspondence to: xy987@uowmail.edu

provided the original work is properly cited.

*DOI: http://dx.doi.org/10.5772/intechopen.91988*

The authors declare no conflict of interest.

**Conflict of interest**

*Coal Burst: A State of the Art on Mechanism and Prevention from Energy Aspect DOI: http://dx.doi.org/10.5772/intechopen.91988*

### **Conflict of interest**

*Mining Techniques - Past, Present and Future*

**4.2 Monitoring**

failure in the monitoring region.

diameter pressure relief drilling.

**4.3 Controlling**

believe that the duration of failure process is the comprehensive reflection of energy accumulation and dissipation characteristics of coal [41]. They propose a dynamic failure time to evaluate coal burst propensity. Based on the correlation analysis of mass data, Qi et al. conclude that uniaxial compression strength of coal is a proper index of coal burst propensity evaluation as well [42]. In 2010, the China Coal Industry Association summarized these four indices as bursting liability indices of coal and published the standard test method of these four indices. Some researchers adopt these four indices to evaluate the burst propensity of rocks as well. It is has been proved by Russian, Poland, and Chinese experience that these four indices are good indicators to define the burst risk of coal seam. Besides, LM Dou et al. combined geological conditions and technical settings of mining together and proposed

comprehensive index method based on the coal burst research in China [16].

Minimizing the safety risk caused by failure of instability rock/coal is an urgent and essential task for underground mines. Similar with the instantaneous failure of other brittle materials such as rock, concrete, and metal, the coal burst process is always associated with the release of rich geophysical signals including acoustic emission (AE) [43], micro-seismic [32] and electromagnetic radiation [44]. It is demonstrated by decades of research and in-field application that micro-seismic monitoring technology has a promising ability to locate potentially violent rock failure. Micro-seismic monitoring is a passive observation of very small-scale earthquakes that occur in the underground as a result of human activities such as mining, hydraulic fracturing, and underground gas storage. The phenomenon that stressed rock can release micro-level signal was discovered by two researchers of U.S. Bureau of Mines, Obert and Duvall, in 1938 [32, 34]. In the early 1960s, South African researchers developed a 16-channel micro-seismic system with positioning function for rock burst monitoring in gold mines [34]. In 1970, under the sponsorship of the U.S. Bureau of Mines, the Pennsylvania State Rock Mechanics Laboratory conducted a research project to investigate the application of micro-seismic techniques to coal mine safety [45]. Through decades' study of underground micro-seismic for mining operation, micro-seismic system has been a basic and valuable monitoring tool for metal and coal mines worldwide. It provides a continuous and real-time 4D (three dimension location and time) record of seismicity associated with rock

The widely used coal burst controlling methods include provocative blasting, long-term water infusion, hydro-fracturing, de-stress drilling, and protective seam mining [46]. Dou et al. proposed the intensity weakening theory to guide the coal burst control from the aspect of energy [16]. Based on the energy aspects, the key to coal burst prevention are: (1) softening coal by changing the physical and mechanical properties of coal. The burst tendency or burst scale of soft coal will be mitigated as the energy storage ability of coal has been reduced. The main methods of coal body softening are blasting and water infusion. (2) Transferring stress to deep regions and reducing the stress level of coal, which can reduce energy storage as well. The main methods are pressure relief blasting, roof pre-splitting blasting, roof cutting blasting, protection seam mining, hydraulic roof fracturing, and large diameter pressure relief drilling. (3) Releasing energy by artificially induced coal burst under low stress level. The main methods are pressure relief blasting and large

**160**

The authors declare no conflict of interest.

#### **Author details**

Xiaohan Yang School of Civil, Mining and Environmental Engineering, University of Wollongong, NSW, Australia

\*Address all correspondence to: xy987@uowmail.edu

© 2020 The Author(s). Licensee IntechOpen. This chapter is distributed under the terms of the Creative Commons Attribution License (http://creativecommons.org/licenses/ by/3.0), which permits unrestricted use, distribution, and reproduction in any medium, provided the original work is properly cited.

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[17] Yang XH, Ren T, Alex R, He XQ, Tan LH. Analysis of energy accumulation and dissipation of coal bursts. Energies. 2018;**11**:1816-1827

[18] Agapito JFT, Goodrich RR. Five stress factors conducive to bursts in Utah, USA, Coal Mines. In: 9th ISRM Congress, International Society for Rock Mechanics. 1999

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Long XTZ, Li ZH. Rock burst tendency of coal-rock combinations sample. Journal of Mining and Safety Engineering. 2006;**23**:43-46

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[20] Zhang CG, Ismet C, Bruce H, Ward CR. Assessing coal burst phenomena in mining and insights into directions for future research. International Journal of Coal Geology. 2017;**179**:28-44

[21] Iannacchione AT, Tadolini SC. Occurrence, predication, and control of coal burst events in The US. International Journal of Mining Science and Technology. 2016;**26**:39-46

[22] Chen XH, Li WQ, Yan XY. Analysis on rock burst danger when fullymechanized caving coal face passed fault with deep mining. Safety Science. 2012;**50**:645-648

[23] Mark C. Coal bursts that occur during development: A rock mechanics enigma. International Journal of Mining Science and Technology. 2017;**28**(1):35-42

[24] Jiang YD, Song HH, Ma ZQ, Ma BJ, Gao LT. Optimization research on the width of narrow coal pillar along goaf tunnel in tectonic stress zone. Journal of China Coal Society. 2018;**43**:319-326

[25] Dou LM, Lu CP, Zong M, Long XTZ, Li ZH. Rock burst tendency of coal-rock combinations sample. Journal of Mining and Safety Engineering. 2006;**23**:43-46

[26] Huang BX, Liu JW. The effect of loading rate on the behavior of samples composed of coal and rock. International Journal of Rock Mechanics and Mining Sciences. 2013;**61**:23-30

[27] Vardar O, Tahmasebinia F, Zhang C, Canbulat I, Saydam S. A review of uncontrolled pillar failures. Procedia Engineering. 2017;**191**:631-637

[28] Whyatt J. Dynamic failure in deep coal: Recent trends and a path forward. In: Singh SP, editor. The 27th International Conference on Ground Control in Mining, Morgantown. 2008 [29] Leśniak A, Isakow Z. Space–time clustering of seismic events and hazard assessment in the Zabrze-Bielszowice coal mine, Poland. International Journal of Rock Mechanics and Mining Sciences. 2009;**46**:918-928

[30] Li ZL, He XQ, Dou LM, Wang GF. Rockburst occurrences and microseismicity in a longwall panel experiencing frequent rockbursts. Geosciences Journal. 2018:1-17

[31] Hallo M. Microseismic surface monitoring network design-sensitivity and accuracy. In: 74th EAGE Conference and Exhibition Incorporating EUROPEC 2012, Copenhagen. 2012

[32] Ge MC. Efficient mine microseismic monitoring. International Journal of Coal Geology. 2005;**64**:44-56

[33] Amitrano D, Arattano M, Chiarle M, Mortara G, Occhiena C, Pirulli M, et al. Microseismic activity analysis for the study of the rupture mechanisms in unstable rock masses. Natural Hazards and Earth System Sciences. 2010;**10**: 831-841

[34] Ge MC. Microseismic monitoring in mines. In: Bruce J, editor. Extracting the Science: A Century of Mining Research, Society for Mining, Metallurgy, and Exploration. Collorado, USA: Littleton; 2010

[35] Li HT, Jiang CX, Jiang YD, Wang HW, Liu HB. Mechanism behavior and mechanism analysis of coal sample based on loading rate effect. Journal of China University of Mining and Tecgnology. 2015;**44**:430-436

[36] Jiang YD, Pan YS, Jiang FX, Dou LM, Ju Y. State of the art review on mechanism and prevention of coal bumps in China. Journal of China Coal Society. 2014;**39**:205-213

[37] Dou LM, He J, Cao AY, Gong SY, Cai W. Rock burst prevention methods based on theory of dynamic and

**162**

*Mining Techniques - Past, Present and Future*

[1] Yang XH, Ren T, He XQ , Tan LH. A review of energy sources of coal burst in Australian coal mines. In: Naj A, Bob K, editors. 2019 Coal Operators Conference, University of Wollongong, characteristics of the burst rock. Scientia Geologica Sinica. 1991:193-200

[11] Dou LM, He J, Cao AY, Cai W, Li ZL, Zhang JW. Mechanism and prevention methods discussion on coal mine rock burst induced by dynamic load. In: High Level Academic Forum of Fifty Anniversary of China National

Coal Association, Beijing. 2012

Abstracts. 1969;**6**:323-330

[12] Bieniawski ZT, Denkhaus HG, Vogler UW. Failure of fractured rock. International Journal of Rock Mechanics & Mining Sciences & Geomechanics

[13] Frid V. Electromagnetic radiation method water-infusion control in rockburst-prone strata. Journal of Applied Geophysics. 2000;**43**:5-13

[14] Dou LM, Zhao CG, Yang SG,

[15] Mark C. Coal bursts in the deep longwall mines of the United States. International Journal of Coal Science &

[16] Dou LM, He XQ. Theory and Technology of Rock Burst Prevention. 1st ed. Xuzhou: China University of Mining and Technology Press; 2001

[17] Yang XH, Ren T, Alex R, He XQ, Tan LH. Analysis of energy accumulation and dissipation of coal bursts. Energies. 2018;**11**:1816-1827

Mechanics. 1999

Industry; 2016

[18] Agapito JFT, Goodrich RR. Five stress factors conducive to bursts in Utah, USA, Coal Mines. In: 9th ISRM Congress, International Society for Rock

[19] Mine Safety. IIR16-05 Austar Coal Burst. Australia: NSW Department of

Technology. 2016;**3**:1-9

Press; 2006

Wu XR. Prevention and Control of Rock Burst in Coal Mine. Xuzhou: China University of Mining and Technology

[2] Pan YS. Study on Rockburst Initiation and Failure Propagation. Doctoral Dissertation; Tsinghua University; 1999

[3] Wang WX, Pan CL, Feng T. Fountain rockburst and inductive rockburst. Journal of Central South University.

[4] Braeuner G. Rockbursts in Coal Mines and their Prevention. Boca Raton:

[5] Zhou XJ, Xian XF. Research advance on rockburst theory and its engineering application in collieries. Journal of Chongqing University (Nature Science

[6] Bruce H, Jim G. A review of the geomechanics aspects of a double fatality coal burst at Austar colliery in NSW, Australia in April 2014.

International Journal of Mining Science

[7] Justine C, Jan N. Coalburst causes and mechanisms. In: Coal Operators

[8] Obert L, Duvall W. Use of subaudible noises for the prediction of rock bursts. In: Technical Report Archive & Image

Wollongong; 2019

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**164**

### *Edited by Abhay Soni*

Mining techniques have evolved over time, culminating in the well-defined field of "mining science," which encompasses aspects such as engineering, chemistry, physics, technology, and management, among others. This book explains how mining techniques can be handled and improved further to make mining practices far more productive, safe, and eco-friendly. It is a useful resource for researchers, students, policy formulators, and decision-makers in different areas of mining and engineering.

Published in London, UK © 2021 IntechOpen © Peter\_Virag / iStock

Mining Techniques - Past, Present and Future

Mining Techniques

Past, Present and Future

*Edited by Abhay Soni*